text stringlengths 20 1.01M | url stringlengths 14 1.25k | dump stringlengths 9 15 ⌀ | lang stringclasses 4
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Integrate with Facebook to help you build engaging social apps and get more installs.
Check out this quick tour to find the best demos and examples for you, and to see how the Felgo SDK can help you to develop your next app or game!
Integrate with Facebook to help you build engaging social apps and get more installs.
Allow people using your app to publish content to Facebook. When people of Facebook engage with these posts they are directed to your app or your app's App Store page (if they don't have your app installed).
Let your users easily sign in to your app with their Facebook account. If they have already signed in, they don't have to reenter their username and password again.
The Graph API is the primary way to get data in and out of Facebook's social graph. You can use it to query data, post new stories, upload photos and a lot of other use cases.
Allow your users using your app to publish from your app to Facebook. When people of Facebook engage with these posts they are directed to your app or your app's App Store page (if they don't have your app installed).
You can use the Facebook item to retrieve information about a logged in user and to send read and publish graph requests.
Facebook can help your apps & games become more popular and engaging for your users.
Retrieve information about your users and their friends
Integrating the Facebook plugin allows your app to get additional information about your users. You can greet them by their first name or save time during a registration process by prefilling the gender or birth date from Facebook.
Reach other people by posting to a User's or App Wall
You can post a message directly to a user's feed, to one of the user's friend's or to your page's feed by using Facebook::postGraphRequest(). See the Direct Wall Post Example for implementation details.
Add a Log In to your App
The Facebook plugin integration allows people to log into your app quickly without having to remember another username and password and therefore improves your conversion. A Facebook login also allows to login to the same account across different devices and platforms.
Like a Facebook Page or App
There is no default interaction of liking Facebook app pages, as initially opening a session with openSession() already registers a user with your app. Instead of liking an app, you can like specific objects in your app or game, e.g. different levels, achievements, etc. These are called stories that follow the "actor verb noun" syntax, for example "Karen likes MyGame Level 1".
Additionally to the use cases for your apps you can also make advantage of the Facebook plugin when integrating in your games. In the following sections, the most popular use-cases are explained:
Brag About Game Progress
It's common practice in mobile games that whenever a player reaches a goal or makes some progress in a game, he can post about it in his news feed (also referred to as his "wall"). The user will then enter a personal message or opinion about something noteworthy that happened in the game, which makes it more interesting for others in comparison to pre-filled text. The message can either be posted on the logged in user's wall, to a friend's wall or to the wall of the Facebook page of your game.
Scores
In addition to other scoring services like Felgo Game Network Facebook can also be used for storing scores online. This has the advantage that friends of your users can compare their results and thus get motivated to beat
the other's highscores. Have a look at the Score API to see how to integrate scores by using Facebook::postGraphRequest() and getGraphRequest() with the
"me/scores" request type.
Invite Friends
Allowing to invite the player's friends to your game can be a major driver of game downloads. You can filter the friends who did not download your game yet to avoid multiple requests.
Match-Making with other Facebook Users
If you are using any kind of multiplayer functionality like in turn-based games, you can select other players that are already using your game as gaming partners. These do not necessarily need to be friends with the logged in player.
Further Game-Related Functionality Available with Facebook
For more information on gaming-related functionality that Facebook offers have a look at.
The Facebook plugin supports single sign-on (SSO) which makes it convenient for your users to login to Facebook.
The SSO feature allows your users, to login to your app with their existing Facebook account and therefore simplifies the login process.
There are 3 different scenarios for SSO:
In scenario 1 and 2 the user is asked to give your app the requested permission defined in Facebook::readPermissions and Facebook::publishPermissions without the need of entering his Facebook credentials beforehand. These are the most convenient methods for your users.
In all other cases the plugin open a web view as a fallback which asks your users to enter their login credentials and afterwards to grant the app permissions. Once the users are logged in, the credentials are stored as cookies in the web view and it's not required to enter the credentials again for every continuing openSession() calls.
So in both cases, either native Facebook integration or not, the user only needs to log in once for the lifetime of your application.
Since the login credentials are stored as explained before, changing to another Facebook user which is often needed during development requires some additional steps. If the native Facebook app is installed, you need to logout in the native app to be able to login with another user after the next call of openSession(). If you are testing on iOS and do not have the native Facebook app installed, open Safari and also logout there, because the login credentials are stored as cookies. Also make sure, to log out your Facebook account in the iOS Settings in the Facebook section, where the application-wide login credentials are stored.
The following sections describe the steps required for adding Facebook connection to your game.
Here is a quick overview of the steps needed:
Go to and create a new Facebook app. On the dashboard, add Facebook Login by clicking "Set Up". You can skip the quickstart guide, as the Facebook item already handles these steps.
In the app settings you now see your App ID and App Secret, which you will need in the next step. You should also create a Facebook canvas page, which is shown when users click on the Facebook graph stories in their web
browser. If this html page is for example hosted on, add
felgo.com to the App Domains. The
following screenshot shows the settings of a test application, where we set the canvas url to the one of ChickenOutbreak Demo.
To test your app or game on iOS & Android, add these platforms in the Facebook settings. Enter your app identifier you have set in the config.json
file as the iOS
Bundle ID and Android
Package Name. The
Class Name should be set to
net.vplay.helper.VPlayActivity, which is also set in your
android/AndroidManifest.xml configuration. For Android, also add the Key Hashes for each of your used signing certificates. See Facebook::printAndroidKeyHash() on how to get your key hashes. The following screenshot shows example settings to configure your Facebook app for iOS & Android.
Note: The Deep Linking setting will launch your app or game when users click on a message in their timeline. We recommend to enable it, as it brings players back to your app or game.
After creating the Facebook app, add the Facebook component to your main qml file. The following example shows the Facebook additions: add the
import Felgo 3.0 statement, the
Facebook component with your Facebook::appId and set the facebookItem property to the id of your Facebook item.
import Felgo 3.0 import QtQuick 2.0 GameWindow { Facebook { id: facebook // this is the Facebook App Id received from the settings in developers.facebook.com/apps appId: "569422866476284" // the permissions define what your app is allowed to access from the user readPermissions: [ "public_profile", "email", "user_friends" ] publishPermissions: [ "publish_actions" ] } // the Scene will follow here }// GameWindow
Facebook requires you to submit your app for review. They need to approve each type of user data item your app wants to access. These items are, for example, the user's friends, posts, or liked pages. In the developer dashboard, click "App Review".
Then click "Start a Submission". Here you can add the privacy concerning permissions your app or game uses.
For example, to allow interacting with the user's friends, you need the friends list permission. For this, check the
user_friends item from the list.
Facebook requires a description of how your app or game uses the permissions. Click "details" and describe how to use Facebook login and the permission step by step.
You will also need to upload a video how the Facebook login works in your app or game. You can use a screen recording tool to create a video similar to this:
Then you can submit your app for review. It usually takes between 24 hours to 14 days until your app gets approved. Once the app was approved, you can use the Facebook item, for example using Facebook::openSession().
To try the plugin or see an integration example have a look at the Felgo Plugin Demo app.
Please also have a look at our plugin example project on GitHub:.
The Facebook item can be used like any other QML item. Here is a simple example of how to add a simple Facebook integration to your project:
import Felgo 3.0 Facebook { appId: "xxxxxxxxxxxxxxxx" readPermissions: [ "public_profile", "email", "user_friends" ] publishPermissions: ["publish_actions"] Component.onCompleted: { openSession() } }
Note: The
user_friends permission only allows to get a list of friends that also use your app and are connected to Facebook.
Before any Facebook interaction you have to open a valid session first. The following example opens a session at app startup and prints information about session state changes to the console:
Facebook { id: facebook appId: "YOUR_APP_ID" onSessionStateChanged: { if (sessionState === Facebook.SessionOpened) { console.debug("Session opened."); // Session opened, you can now perform any Facebook actions with the plugin! } else if (sessionState === Facebook.SessionOpening) { console.debug("Session opening..."); } else if (sessionState === Facebook.SessionClosed) { console.debug("Session closed."); } else if (sessionState === Facebook.SessionFailed) { console.debug("Session failed."); } else if (sessionState === Facebook.SessionPermissionDenied) { console.debug("User denied requested permissions."); } } Component.onCompleted: { facebook.openSession(); } }
You can send a pre-defined message with a defined message text with the Facebook::postGraphRequest() method. The following example sends a message to the own wall, without any further parameters like a link, description or caption. These parameters are also available for this request though, so you can set them if required.
facebook.postGraphRequest("me/feed", { "message" : "Hello me!" })
You can also provide a
to parameter to post on a friend's wall or on a Facebook page. Please keep in mind that posting to a Facebook page is only possible when the logged in user liked the page before.
Posting directly to walls usually has a weaker growth effect than adding a personal message, because the friends' engagement and interest in personal messages is higher. However, you could still open a native input dialog and send the message with the above call afterwards. This has the advantage that the user needs not leave the app and no Facebook dialog is shown. Also, the native Facebook app is not launched and the user stays in your game.
Note: Posting on the user wall requires the
publish_actions permissions. At the first post request, the user must allow the publish permission for your app. Facebook apps that use this permission also
require a review by Facebook before the app can be used by people other than you as the app developer. Startup or Business license. To activate plugins and enable their full functionality it is required to create a license key. You can create such a key for your application using the license creation page.
This is how it works:
To use the Facebook plugin you need to add the platform-specific native libraries to your project, described here:
Add the following lines of code to your
.pro file:
FELGO_PLUGINS += facebook
FBSDKCoreKit.frameworkand
FBSDKLoginKit.frameworkfrom the
iossubfolder to a sub-folder called
ioswithin your project directory.
.profile:
ios { LIBS += -framework Accelerate }
Note: Adding the
Accelerate framework manually is a hotfix that only applies to Felgo 3.3.0.
Project-Info.plistfile within the
iossubfolder of your project and add the following lines of code:
<key>CFBundleURLTypes</key> <array> <dict> <key>CFBundleURLSchemes</key> <array> <string>fb{your-app-id}</string> </array> </dict> </array> <key>LSApplicationQueriesSchemes</key> <array> <string>fbapi</string> <string>fb-messenger-api</string> <string>fbauth2</string> <string>fbshareextension</string> </array>
right before the closing tags:
</dict> </plist>
Make sure to replace
{your-app-id}with your Facebook app's app ID.
build.gradlefile and add the following lines to the dependencies block:
dependencies { implementation 'net.vplay.plugins:plugin-facebook:3.+' }
Note: If you did not create your project from any of our latest wizards, make sure that your project uses the Gradle Build System like described here.
In order to use the Facebook plugin within your app you need to create a Facebook app at first.
The Google Play Package Name of your app, found in the AndroidManifest.xml file.
The Class Name as
net.vplay.helper.VPlayActivity (also set in your project's AndroidManifest.xml file).
The Key Hashes for each of your used signing certificates. See Facebook::printAndroidKeyHash() on how to get you key hashes (the hash in the screenshot above is only a sample value!).
Note: Other SDK versions higher than the stated ones might also be working but are not actively tested as of now.
See also Facebook::openSession(), Facebook::sessionState, Facebook::publishPermissions, and Facebook::postGraphRequest(). | https://felgo.com/doc/plugin-facebook/ | CC-MAIN-2021-04 | en | refinedweb |
Calling Go code from Python code
Inspired by the ruby version from @jondot
With the release of Go 1.5 you can now call Go as a shared library in other languages. That little tidbit escaped me the first time I went through the release notes but after having the above ruby article come through my twitter stream I decided I should see if I could quickly get a simple example working between Go and Python.
First you have to write a short go func you want to expose to your python code in a simple file. I wrote one in libadd.go. The comments are critical to include so the functions are exported.
//libadd.go package main import "C" //export add func add(left, right int) int { return left + right } func main() { }
Then you’re going to want to use the new buildmode option in 1.5
go build -buildmode=c-shared -o libadd.so libadd.go
Make sure you’ve upgraded to 1.5.1 while having completely removed any previous version or you’ll get this error
imports runtime: C source files not allowed when not using cgo or SWIG: atomic_amd64x.c defs.c float.c heapdump.c lfstack.c malloc.c mcache.c mcentral.c mem_linux.c mfixalloc.c mgc0.c mheap.c msize.c os_linux.c panic.c parfor.c proc.c runtime.c signal.c signal_amd64x.c signal_unix.c stack.c string.c sys_x86.c vdso_linux_amd64.c
The python code is really short and this is only passing an integer back and forth (more complex string and struct cases are much more challenging).
from ctypes import cdll lib = cdll.LoadLibrary('./libadd.so') print "Loaded go generated SO library" result = lib.add(2, 3) print result
The Ruby post includes a full demo project that shows what the go code would look like that would go fetch several URLs using channels and return results back that project is here and it would probably be worth doing a Python version at some point the illustrates how to do more complex struct and types (this was a very limited post).
Here is some further reading:
the design doc around how this is going to all work
python ctypes and cffi
cgo and how types interplay with the export/import go C package
ctypes documentation
cffi documentation
Perhaps a note to indicate that the “//export add” comment is important!!!
Added, thanks!
Would be better if you post more examples, because I’m adding a factorial function with channel and goroutines inside it, I got error
Traceback (most recent call last):
File “add.py”, line 7, in
resfib = lib.factorial(6)
File “/usr/local/var/pyenv/versions/3.5.0/lib/python3.5/ctypes/__init__.py”, line 360, in __getattr__
func = self.__getitem__(name)
File “/usr/local/var/pyenv/versions/3.5.0/lib/python3.5/ctypes/__init__.py”, line 365, in __getitem__
func = self._FuncPtr((name_or_ordinal, self))
AttributeError: dlsym(0x7fbd9ac1f6f0, factorial): symbol not found | http://savorywatt.com/2015/09/18/calling-go-code-from-python-code/ | CC-MAIN-2019-47 | en | refinedweb |
In which topic/lesson is the
import.math covered? I think I either have a gap in my notes or I just haven’t covered it for some reason.
Import.math topic - where is it?
In which topic/lesson is the
1 Like
It might come up before the Modules unit, but for sure it is also covered in that unit.
It won’t be written like that, at any length.
import math
All imports should occur at the start of the script (top lines).
We access that namespace by module name dot method name:
math.sqrt() PI = math.pi
Note that
math.pi is a constant, not a method, so we do not invoke it.
Python 2.x Mathematical functions
Python 3.x Mathematical functions
2 Likes | https://discuss.codecademy.com/t/import-math-topic-where-is-it/393273 | CC-MAIN-2019-47 | en | refinedweb |
Extract quartiles and extremum values of all columns of a table or all fields of a dataset. More...
#include <vtkComputeQuartiles.h>
Extract quartiles and extremum values of all columns of a table or all fields of a dataset.
vtkComputeQuartiles accepts any vtkDataObject as input and produces a vtkTable data as output. This filter can be used to generate a table to create box plots using a vtkPlotBox instance. The filter internally uses vtkOrderStatistics to compute quartiles.
Definition at line 49 of file vtkComputeQuartiles.h.
Definition at line 53 of file vtkComputeQuartiles 68 of file vtkComputeQuartiles.h. | https://vtk.org/doc/nightly/html/classvtkComputeQuartiles.html | CC-MAIN-2019-47 | en | refinedweb |
Count the characters at the beginning of a string that aren't in a given character set
#include <string.h> size_t strcspn( const char* str, const char* charset );
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
The strcspn() function finds the length of the initial segment of the string pointed to by str that consists entirely of characters not from the string pointed to by charset. The terminating NUL character isn't considered part of str.
The length of the initial segment.
#include <stdio.h> #include <string.h> #include <stdlib.h> int main( void ) { printf( "%d\n", strcspn( "abcbcadef", "cba" ) ); printf( "%d\n", strcspn( "xxxbcadef", "cba" ) ); printf( "%d\n", strcspn( "123456789", "cba" ) ); return EXIT_SUCCESS; }
produces the output:
0 3 9 | http://www.qnx.com/developers/docs/6.6.0.update/com.qnx.doc.neutrino.lib_ref/topic/s/strcspn.html | CC-MAIN-2019-47 | en | refinedweb |
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destluckMember
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About destluck
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Functions
destluck replied to destluck's topic in For Beginners's ForumPerfect tyvm for the fast response :)
Functions
destluck posted a topic in For Beginners's ForumAlright im not to sure if im using the right terms. Anyhow i wanted to know if there is a difference in declaring/initializing a function before the main or after it. - Is it better to declare it before main or after? - What affect does it cause? example: // function delcared before main() char askYesNo1() { ......... } // main function int main() { return 0; } // function delcared after main() char askYesNo2() { .... }
Any idea what i did wrong?
destluck replied to destluck's topic in For Beginners's Forumty
Any idea what i did wrong?
destluck posted a topic in For Beginners's ForumAlright i keep getting 2 errors: Error 1 error LNK2005: _main already defined in main.obj Error 2 error LNK1169: one or more multiply defined symbols found #include<iostream> #include <string> #include <random> #include <time.h> using namespace std; class Player { public: string mFirstName; string mLastName; string mStreetName; string mGender; int mAdress; int mAge; }; Player GetPlayerFromConsole() { Player npc; cout << " New Player Signup: " << endl; cout << "Enter your first name: " << endl; cin >> npc.mFirstName; cout << "Enter your last name: " << endl; cin >> npc.mLastName; cout << " Enter your Age: " << endl; cin >> npc.mAge; cout << "Enter your Gender: " << endl; cin >> npc.mGender; cout << "Enter your Street name: " << endl; cin >> npc.mStreetName; cout << "Enter your adress: " << endl; cin >> npc.mAdress; return npc; } int main() { Player newPlayer; newPlayer = GetPlayerFromConsole(); cout << "New Player info Sheet: " << endl; cout << "Player name: " << newPlayer.mFirstName << endl; cout << newPlayer.mFirstName << endl; cout << "Player Adress: " << newPlayer.mAdress << newPlayer.mStreetName << endl; cout << "Player Gender: " << newPlayer.mGender << endl; cin.get(); system("pause"); return 0; }
- thx for all the great posts and links reading trough them as we speak :) great stuff
- Thx ill give it a read.
- hehe well that's reassuring i think ... when i finish work in 1h ill be able to pinpoint exactly the parts im having issues with. Normally i learn by knowing what specific things are used for in a game based environment.
Keep getting stuck
destluck posted a topic in For Beginners's ForumAlright i have been trying to learn C++ for a while now. Been buying VTM's and books and i always seem to get stuck. Always around the chapters 6-7 and that turns out to be where pointers kick in. Maybe the books i have a just bad or im just a real slow learner. Anyhow i was wondering if anyone could link me a few things i can read or tutorials that helped you learn C++. If anyone is willing to coach me a bit that would be awesome :) . Any info or help is more then welcomed sick of being stuck or not able to wrap my head around some issues i have. DestLuck
Image detection
destluck replied to destluck's topic in General and Gameplay Programmingthis for the info yah basicly the anti-macro screen pops up randomly and shows 3-5 icons. Then you have a list of icons at the bottom and you must click the icons that match. as you can see in this picture there 3 icons/images from the tops have to be matched with the ones on the bottom. [url=][/URL]
Image detection
destluck posted a topic in General and Gameplay Programmingwhat would be the best language/method to get image detection to work? The anti-Macro brings up a screen with images/icons that you need to match. The icons keep changing. this video should show what i need the program to do. ** Link ** *** *** thx for the info
LF Programmer to Hire!
destluck posted a topic in General and Gameplay ProgrammingLooking to Hire a programmer to help me bypass a game issue. You tell me the price for the project and i will pay you for it. Game: Ashen Empires Payment method: Paypal Payment: Milestone this video should show what i need the program to do. [mod edit: link redacted]
- Advertisement | https://www.gamedev.net/profile/221915-destluck/ | CC-MAIN-2019-47 | en | refinedweb |
bwrap (1) - Linux Man Pages
bwrap: container setup utility
NAMEbwrap - container setup utility
SYNOPSIS
- bwrap [OPTION...] [COMMAND]
DESCRIPTION
bwrap
It works by creating a new, completely empty, filesystem namespace where the root is on a tmpfs that is invisible from the host, and which will be automatically cleaned up when the last process exists. You can then use commandline options to construct the root filesystem and process environment for the command to run in the namespace.
By default, bwrap creates a new mount namespace for the sandbox. Optionally it also sets up new user, ipc, pid, network and uts namespaces (but note the user namespace is required if bwrap is not installed setuid root). The application in the sandbox can be made to run with a different UID and GID.
If needed (e.g. when using a PID namespace) bwrap is running a minimal pid 1 process in the sandbox that is responsible for reaping zombies. It also detects when the initial application process (pid 2) dies and reports its exit status back to the original spawner. The pid 1 process exits to clean up the sandbox when there are no other processes in the sandbox left.
OPTIONS
When options are used multiple times, the last option wins, unless otherwise specified.
General options:
- Print help and exit
--version
- Print version
--args FD
- Parse nul-separated arguments from the given file descriptor. This option can be used multiple times to parse options from multiple sources.
Options related to kernel namespaces:
--unshare-user
- Create a new user namespace
--unshare-user-try
- Create a new user namespace if possible else skip it
--unshare-ipc
- Create a new ipc namespace
--unshare-pid
- Create a new pid namespace
--unshare-net
- Create a new network namespace
--unshare-uts
- Create a new uts namespace
--unshare-cgroup
- Create a new cgroup namespace
--unshare-cgroup-try
- Create a new cgroup namespace if possible else skip it
--unshare-all
- Unshare all possible namespaces. Currently equivalent with: --unshare-user-try --unshare-ipc --unshare-pid --unshare-net --unshare-uts --unshare-cgroup-try
--uid UID
- Use a custom user id in the sandbox (requires --unshare-user)
--gid GID
- Use a custom group id in the sandbox (requires --unshare-user)
--hostname HOSTNAME
- Use a custom hostname in the sandbox (requires --unshare-uts)
Options about environment setup:
--chdir DIR
- Change directory to DIR
--setenv VAR VALUE
- Set an environment variable
--unsetenv VAR
- Unset an environment variable
Options for monitoring the sandbox from the outside:
--lock-file DEST
- Take a lock on DEST while the sandbox is running. This option can be used multiple times to take locks on multiple files.
--sync-fd FD
- Keep this file descriptor open while the sandbox is running
Filesystem related options. These are all operations that modify the filesystem directly, or mounts stuff in the filesystem. These are applied in the order they are given as arguments. Any missing parent directories that are required to create a specified destination are automatically created as needed.
--bind SRC DEST
- Bind mount the host path SRC on DEST
--dev-bind SRC DEST
- Bind mount the host path SRC on DEST, allowing device access
--ro-bind SRC DEST
- Bind mount the host path SRC readonly on DEST
--remount-ro DEST
- Remount the path DEST as readonly. It works only on the specified mount point, without changing any other mount point under the specified path
--proc DEST
- Mount procfs on DEST
--dev DEST
- Mount new devtmpfs on DEST
--tmpfs DEST
- Mount new tmpfs on DEST
--mqueue DEST
- Mount new mqueue on DEST
--dir DEST
- Create a directory at DEST
--file FD DEST
- Copy from the file descriptor FD to DEST
--bind-data FD DEST
- Copy from the file descriptor FD to a file which is bind-mounted on DEST
--ro-bind-data FD DEST
- Copy from the file descriptor FD to a file which is bind-mounted readonly on DEST
--symlink SRC DEST
- Create a symlink at DEST with target SRC
Lockdown options:
--seccomp FD
- Load and use seccomp rules from FD. The rules need to be in the form of a compiled eBPF program, as generated by seccomp_export_bpf.
--exec-label LABEL
- Exec Label from the sandbox. On an SELinux system you can specify the SELinux context for the sandbox process(s).
--file-label LABEL
- File label for temporary sandbox content. On an SELinux system you can specify the SELinux context for the sandbox content.
--block-fd FD
- Block the sandbox on reading from FD until some data is available.
--userns-block-fd FD
- Do not initialize the user namespace but wait on FD until it is ready. This allow external processes (like newuidmap/newgidmap) to setup the user namespace before it is used by the sandbox process.
--info-fd FD
- Write information in JSON format about the sandbox to FD.
--new-session
- Create a new terminal session for the sandbox (calls setsid()). This disconnects the sandbox from the controlling terminal which means the sandbox can't for instance inject input into the terminal.
Note: In a general sandbox, if you don't use --new-session, it is recommended to use seccomp to disallow the TIOCSTI ioctl, otherwise the application can feed keyboard input to the terminal.
--die-with-parent
- Ensures child process (COMMAND) dies when bwrap's parent dies. Kills (SIGKILL) all bwrap sandbox processes in sequence from parent to child including COMMAND process when bwrap or bwrap's parent dies. See prctl, PR_SET_PDEATHSIG.
--as-pid-1
- Do not create a process with PID=1 in the sandbox to reap child processes.
--cap-add CAP
- Add the specified capability when running as privileged user. It accepts the special value ALL to add all the permitted caps.
--cap-drop CAP
- Drop the specified capability when running as privileged user. It accepts the special value ALL to drop all the caps. By default no caps are left in the sandboxed process. The --cap-add and --cap-drop options are processed in the order they are specified on the command line. Please be careful to the order they are specified.
ENVIRONMENT
- Used as the cwd in the sandbox if --cwd has not been explicitly specified and the current cwd is not present inside the sandbox. The --setenv option can be used to override the value that is used here.
EXIT STATUS
The bwrap command returns the exit status of the initial application process (pid 2 in the sandbox).
Linux man pages generated by: SysTutorials | https://www.systutorials.com/docs/linux/man/1-bwrap/ | CC-MAIN-2019-47 | en | refinedweb |
General Information
Desktop
Web
Controls and Extensions
Mainteinance Mode
Enterprise and Analytic Tools
Quality Assurance and Productivity
Frameworks and Libraries
Ways to Register a Module
- 5 min to read
In an XAF application, you can use modules declared in the current solution, as well as modules provided by external assemblies. This topic lists the ways you can follow to register a module in your applications.
- Use the Solution Wizard
- Use the Module Designer or Application Designer
-
- Add a Module to the Application Configuration File
Use the Solution Wizard
You can add modules to your application when you create a new XAF solution using the Solution Wizard. To do this, select modules in the Choose Extra Modules step.
Note
The Solution Wizard allows you to add XAF modules from the predefined list. To add a custom module, use other approaches described in this topic.
Use the Module Designer or Application Designer
In an existing XAF solution, start the Application Designer or Module Designer. Drag the required module from the Toolbox to the Designer's Required Modules / Modules section.
Modules that are shipped with XAF (Extra Modules) are available in the DX.19.2: XAF Modules tab of the Toolbox. If the module is custom and is declared in an external assembly, register this module in the Toolbox, so that you can drag it to the Application or Module Designer when required. For details, refer to the How to: Add Items to the Toolbox topic in MSDN.
Add a Module in Code
This section lists several ways you can follow to add a module in code. The general restriction is that the code should be executed before the XafApplication.Setup method is called. In XAF, adding modules is not supported when the Setup method is already executed. You can register modules dynamically depending on a certain condition. For example, you can show a custom form when the application is started and register certain modules depending on a user input.
Register the Module Type Using the ModuleBase.RequiredModuleTypes Property
In an existing module project, you can register an extra module that will be loaded with the current module. In the module constructor (which is declared in the Module.cs (Module.vb) file by default), add the required module to the ModuleBase.RequiredModuleTypes list.
Add the Module Instance to the XafApplication.Modules List
In the application project, you can instantiate the required module and add the module object to the XafApplication.Modules list. For instance, you can do it in the WinApplication/WebApplication descendant constructor.
Pass the Module Assembly Name to the XafApplication.Setup Method
The advantage of this approach is that you do not have to explicitly reference the required module assembly. The assembly will be loaded dynamically by its name when required. There is an overload of the XafApplication.Setup method taking the moduleAssemblies parameter - a string array specifying module assembly names to be loaded. For instance, you can use this method to load modules listed in the application configuration file. Proceed to the next section to see the example.
Add a Module to the Application Configuration File
This approach allows third-parties to plug-in their own modules without recompiling your application. Please note that third-party modules may contain insecure code that overrides security restrictions, modifies your business logic, etc. That is why adding modules from the configuration file is disabled by default. Enable it in the trusted environment only.
In a WinForms application, edit the Program.cs (Program.vb) file.
static class Program { static void Main() { //... MySolutionWindowsFormsApplication winApplication = new MySolutionWindowsFormsApplication(); //... winApplication.Setup("MySolution", winApplication.ConnectionString, ConfigurationManager.AppSettings["Modules"].Split(';')); winApplication.Start(); //... } //... }
In an ASP.NET application, edit the Global.asax.cs (Global.asax.vb) file.
public class Global : System.Web.HttpApplication { protected void Session_Start(object sender, EventArgs e) { WebApplication.SetInstance(Session, new MySolutionWebApplication()); //... WebApplication.Instance.Setup("MySolution", WebApplication.Instance.ConnectionString, ConfigurationManager.AppSettings["Modules"].Split(';')); WebApplication.Instance.Start(); } //... }
You can now add a module assembly name to the modules list in the application configuration file (App.config and/or Web.config).
<configuration> <appSettings> <add key="Modules" value="MySolution.MyCustomModule" /> </appSettings> </configuration>
Here, MySolution.MyCustomModule is a custom module assembly name.
Note
If the module to be added includes the Entity Framework DbContext, register the EFObjectSpace provider for it at the module level. Then, override the ModuleBase.Setup method and handle the XafApplication.CreateCustomObjectSpaceProvider event. | https://docs.devexpress.com/eXpressAppFramework/118047/concepts/application-solution-components/ways-to-register-a-module | CC-MAIN-2019-47 | en | refinedweb |
Chages Net::GPSD 2010-06-01 0.39 2010-01-02 0.38 - GPS-PRN moved to GPS-OID 2009-01-18 0.37 - Added socket caching (Bug RT 38406) 2007-09-15 0.36 - Added set to Net::GPSD->host and Net::GPSD->port methods - Updated Net::GPSD->speed_knots method - Added example-googleearth.cgi 2006-01-15 0.35 - Added Net::GPSD->speed_knots method 2006-01-03 0.34 - Added extensions to example scripts 2006-12-17 0.31 - added oid satellite capability from GPS::PRN 2006-12-02 0.30 - cleaned 2006-12-01 0.29 - updated test near function - Distance function from Geo::Inverse 2006-11-30 0.28 - documentation - getsatellite method supports wantarray - moved PI from sub to constant - moved stuff around everywhere but no real changes 2006-10-28 0.27 - changed track formula to use Geo:Forward. 2006-10-11 0.26 - Error in Net::GPSD->track $||$*$ -> ($||$)*$; 2006-06-19 0.25 - Change dependancy on S[0] first and then O[14]||M[0] to get fix. - Added examples 2006-06-14 0.24 - Change dependancy on O[14](new) and M[0] vice S[0] to get fix. 2006-06-11 0.23 - added logic to handle O=? watcher no fix 2006-06-08 0.22 - modified subscribe method to use gpsd watcher mode vs. poll mode 2006-06-08 0.21 - updated example-tracker-text 2006-06-08 0.20 - Scrapped development concerning Math::Bezier 2006-04-23 0.19 - added example-tracker-text - added Point latlon method - added wantarray capability to commands method - changed a connection error print from stdout to a warn on stderr 2006-04-08 0.18 - shift() warning - fixed test errors 2006-04-08 0.17 - Forgot to update versions - Updated the CHANGES file 2006-04-08 0.16 - Error in track function > # Heading is measured clockwise from the North. The angles for the math > # sin and cos formulas are measured anti-clockwise from the East. So, > # in order to get this correct, we have to shift sin and cos the 90 > # degrees to cos and -sin. The anti-clockwise/clockwise change flips > # the sign on the sin back to positive. > my $distance_lat_meters=$distance_meters * cos(deg2rad $p1->heading); > my $distance_lon_meters=$distance_meters * sin(deg2rad $p1->heading); - Added deg2rad function - Added Point->latlon function 2006-04-01 0.15 - Updated pod for GPSD.pm mostly just the examples are linked 2006-03-29 0.14 - Renamed GPS::gpsd to Net::GPSD 2006-03-22 0.13 - Internel version numbers were wrong 2006-03-22 0.12 - Error in point copy for initialization < $self->{$_}=$data->{$_}; > $self->{$_}=[@{$data->{$_}}]; 2006-03-21 0.11 - simplified GPSD default_handler - pods for ./bin/ examples - hopefully fixed META.yml error 2006-03-19 0.10 - moved examples to ./bin/ folder 2006-03-17 0.09 - fixed 1 warning 2006-03-15 0.08 - CPAN changes. Now automakes with CPAN! - moved Report::http under gpsd namespace - moved modules to the lib folder - renamed tgz file to GPS-gpsd-X.XX.tgz format 2006-03-11 0.07 - made a bunch of changes to the distance calculations - Fixes error in in the parse routine < $data{$1}=[split(/ /, $2)]; > $data{$1}=[split(/\s+/, $2)]; - Updates to CPAN install capability - Updates to documentation - Update to the subscribe interface 2006-02-23 0.06 - No user interface changes - Updates the pod documentation so that it displays better on CPAN. - Moved code from GPS::gpsd::Satellite->list to GPS::gpsd->getsatellitelist. - Documentation, Documentation, Documentation... 2006-02-22 0.05 - Heavy user interface changes - Modified a few interface names to meet my tastes register->subscribe - Documentation, Documentation, Documentation... 2006-02-21 0.04 - Heavy user interface changes - First CPAN Documentation Begins - Added satellite object interface 2006-02-19 0.03 - Heavy user interface changes - Added Point object interface 2006-02-?? 0.02 - Heavy user interface changes 2006-02-?? 0.01 - Original version on CPAN. | http://web-stage.metacpan.org/changes/distribution/Net-GPSD | CC-MAIN-2019-47 | en | refinedweb |
Unlock the range of process address space already allocated
#include <sys/mman.h> int munlock(const void * addr, size_t len);
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
The munlock() function unlocks the specified range of address space, which was locked by a call to mlock() or mlockall(). Memory locks don't stack; the memory is unlocked, no matter how many times the same process locked it.
The munlock() function doesn't affect those pages that have been mapped and locked into the address spaces of other processes. | http://www.qnx.com/developers/docs/6.6.0.update/com.qnx.doc.neutrino.lib_ref/topic/m/munlock.html | CC-MAIN-2019-47 | en | refinedweb |
Can you help me make a Hotkey please?
I have a stack with 2 renders and I would like to be able to toggle the Composite Mode -> Difference on and off.
I'm hoping ctrl+d will toggle between OVER and DIFFERENCE.
thanks
Hi Nick!
Hotkeys can be added through packages that run either python or mu code. Toggling the composite mode is just a single property, but to discover the property; you can either check out the reference manual for the property ( ) or my favorite way is the save out the session file before I toggle the parameter and afterward and view the difference.
Here I've taken both over and difference and saved each out as a session file and brought them into my diff viewer of choice (meld).
You can see the only change is what we would expect, a change from over to difference; contained within the "stack" node under the component block "composite"
We now know the property we need to change is (in this case) default_stack.composite.type
To build the package, the minimal example in both mu and python is described here:
If we take that package as a framework, we can build a function to call that swap between the two states with something like this:
import rv.commands
property_name ="@RVStack.composite.type"
comp_types = rv.commands.getStringProperty(property_name)
if comp_types:
if comp_types[0] == "over":
rv.commands.setStringProperty(property_name, ["difference"])
else:
rv.commands.setStringProperty(property_name, ["over"])
That will look for the first RV stack's composite type.
If it finds one it will check to see if it is already set to "over".
If it is set to over, then set it instead to difference; otherwise, set it to over.
Once we integrate it into a simple package, we have something that looks like this:
from rv import commands, rvtypes, runtime
from rv.commands import NeutralMenuState
class ToggleCompositeMode(rvtypes.MinorMode):
def __init__(self):
rvtypes.MinorMode.__init__(self)
self.init("ToggleCompositeMode", [("key-down--control--d", self.toggleMode, "Toggle Composite Mode")], None, None)
def toggleMode(self, e=None):
property_name ="@RVStack.composite.type"
comp_types = commands.getStringProperty(property_name)
if comp_types:
if comp_types[0] == "over":
commands.setStringProperty(property_name, ["difference"])
else:
commands.setStringProperty(property_name, ["over"])
def createMode():
return ToggleCompositeMode()
I've packaged the above into an .rvpkg file as an example, but all that requires is writing a PACKAGE file and zipping the two up; that format is described here:
You can download the example package I wrote for this post here:
Nick, It appears that I linked the wrong package; the above one is a similar (though more complex example) of setting the annotation undo to a keyboard shortcut.
The proper link is here:
Hiya,
Thanks for your help :D
I'm not able to download the package as I dont have a Box@Autodesk account. The link takes me to a Login page.
Would it be possible to email me ? <email removed>
also.. i'm using 6.2.7
I've just opened up the permissions; but I'll also email it to you.
that is awesome! works well, Thanks.
Is the createMode() function required by RV? Can this be renamed per package, or does it have to be in each package?
Yes, the createMode() is called by RV and needs to exist for the mode to be registered by the package manager. It needs to be in each package, but in most cases it is that simple boilerplate code, ie:
def createMode():
return ToggleCompositeMode()
If you were to get fancy and need to set up things at mode definition time (or profiling, etc) you could do it there, but in most cases the implementation will be as simple as above.
Hi, I know this is a super old thread but is there any way to currently access the examples Michael Kessler mentioned? Thanks in advance!! | https://support.shotgunsoftware.com/hc/zh-cn/community/posts/115003946913-Need-Help-Making-a-Hotkey?page=1 | CC-MAIN-2019-47 | en | refinedweb |
#include <deal.II/lac/block_sparsity_pattern.h>
This is the base class for block versions of the sparsity pattern and dynamic sparsity pattern classes. It has not much functionality, but only administrates an array of sparsity pattern objects and delegates work to them. It has mostly the same interface as has the SparsityPattern, and DynamicSparsityPattern, and simply transforms calls to its member functions to calls to the respective member functions of the member sparsity patterns.
The largest difference between the SparsityPattern and DynamicSparsityPattern classes and this class is that mostly, the matrices have different properties and you will want to work on the blocks making up the matrix rather than the whole matrix. You can access the different blocks using the
block(row,col) function.
Attention: this object is not automatically notified if the size of one of its subobjects' size is changed. After you initialize the sizes of the subobjects, you will therefore have to call the
collect_sizes() function of this class! Note that, of course, all sub-matrices in a (block-)row have to have the same number of rows, and that all sub-matrices in a (block-)column have to have the same number of columns.
You will in general not want to use this class, but one of the derived classes.
Definition at line 1905 of file affine_constraints.h.
Declare type for container size.
Definition at line 85 of file block_sparsity_pattern.h.
Initialize the matrix empty, that is with no memory allocated. This is useful if you want such objects as member variables in other classes. You can make the structure usable by calling the reinit() function.
Definition at line 25 of file block_sparsity_pattern.cc.
Initialize the matrix with the given number of block rows and columns. The blocks themselves are still empty, and you have to call collect_sizes() after you assign them sizes.
Definition at line 33 of file block_sparsity_pattern.cc.
Copy constructor. This constructor is only allowed to be called if the sparsity pattern to be copied is empty, i.e. there are no block allocated at present. This is for the same reason as for the SparsityPattern, see there for the details.
Definition at line 45 of file block_sparsity_pattern.cc.
Destructor.
Definition at line 62 of file block_sparsity_pattern.cc.
Resize the matrix, by setting the number of block rows and columns. This deletes all blocks and replaces them with uninitialized ones, i.e. ones for which also the sizes are not yet set. You have to do that by calling the reinit() functions of the blocks themselves. Do not forget to call collect_sizes() after that on this object.
The reason that you have to set sizes of the blocks yourself is that the sizes may be varying, the maximum number of elements per row may be varying, etc. It is simpler not to reproduce the interface of the SparsityPattern class here but rather let the user call whatever function she desires.
Definition at line 77 of file block_sparsity_pattern.cc.
Copy operator. For this the same holds as for the copy constructor: it is declared, defined and fine to be called, but the latter only for empty objects.
Definition at line 111 of file block_sparsity_pattern.cc.
This function collects the sizes of the sub-objects and stores them in internal arrays, in order to be able to relay global indices into the matrix to indices into the subobjects. You must call this function each time after you have changed the size of the sub-objects.
Definition at line 129 of file block_sparsity_pattern.cc.
Access the block with the given coordinates.
Definition at line 767 of file block_sparsity_pattern.h.
Access the block with the given coordinates. Version for constant objects.
Definition at line 779 of file block_sparsity_pattern.h.
Grant access to the object describing the distribution of row indices to the individual blocks.
Definition at line 792 of file block_sparsity_pattern.h.
Grant access to the object describing the distribution of column indices to the individual blocks.
Definition at line 801 of file block_sparsity_pattern.h.
This function compresses the sparsity structures that this object represents. It simply calls
compress for all sub-objects.
Definition at line 168 of file block_sparsity_pattern.cc.
Return the number of blocks in a column.
Definition at line 962 of file block_sparsity_pattern.h.
Return the number of blocks in a row.
Definition at line 953 of file block_sparsity_pattern.h.
Return whether the object is empty. It is empty if no memory is allocated, which is the same as that both dimensions are zero. This function is just the concatenation of the respective call to all sub- matrices.
Definition at line 179 of file block_sparsity_pattern.cc.
Return the maximum number of entries per row. It returns the maximal number of entries per row accumulated over all blocks in a row, and the maximum over all rows.
Definition at line 192 of file block_sparsity_pattern.cc.
Add a nonzero entry to the matrix. This function may only be called for non-compressed sparsity patterns.
If the entry already exists, nothing bad happens.
This function simply finds out to which block
(i,j) belongs and then relays to that block.
Definition at line 810 of file block_sparsity_pattern.h.
Add several nonzero entries to the specified matrix row. This function may only be called for non-compressed sparsity patterns.
If some of the entries already exist, nothing bad happens.
This function simply finds out to which blocks
(row,col) for
col in the iterator range belong and then relays to those blocks.
Definition at line 829 of file block_sparsity_pattern.h.
Return number of rows of this matrix, which equals the dimension of the image space. It is the sum of rows of the (block-)rows of sub-matrices.
Definition at line 211 of file block_sparsity_pattern.cc.
Return number of columns of this matrix, which equals the dimension of the range space. It is the sum of columns of the (block-)columns of sub- matrices.
Definition at line 225 of file block_sparsity_pattern.cc.
Check if a value at a certain position may be non-zero.
Definition at line 917 of file block_sparsity_pattern.h.
Number of entries in a specific row, added up over all the blocks that form this row.
Definition at line 935 of file block_sparsity_pattern.h..
In the present context, it is the sum of the values as returned by the sub-objects.
Definition at line 239 of file block_sparsity_pattern.cc.
Print the sparsity of the matrix. The output consists of one line per row of the format
[i,j1,j2,j3,...]. i is the row number and jn are the allocated columns in this row.
Definition at line 252 of file block_sparsity_pattern.cc.
Print the sparsity of the matrix in a format that
gnuplot understands and which can be used to plot the sparsity pattern in a graphical way. This is the same functionality implemented for usual sparsity patterns, see SparsityPattern.
Definition at line 306 of file block_sparsity_pattern.cc.
Typedef for the type used to describe sparse matrices that consist of multiple blocks.
Definition at line 385 of file block_sparsity_pattern.h.
Define a value which is used to indicate that a certain value in the
colnums array is unused, i.e. does not represent a certain column number index.
This value is only an alias to the respective value of the SparsityPattern class.
Definition at line 95 of file block_sparsity_pattern.h.
Number of block rows.
Definition at line 342 of file block_sparsity_pattern.h.
Number of block columns.
Definition at line 347 of file block_sparsity_pattern.h.
Array of sparsity patterns.
Definition at line 355 of file block_sparsity_pattern.h.
Object storing and managing the transformation of row indices to indices of the sub-objects.
Definition at line 361 of file block_sparsity_pattern.h.
Object storing and managing the transformation of column indices to indices of the sub-objects.
Definition at line 367 of file block_sparsity_pattern.h.
Temporary vector for counting the elements written into the individual blocks when doing a collective add or set.
Definition at line 374 of file block_sparsity_pattern.h.
Temporary vector for column indices on each block when writing local to global data on each sparse matrix.
Definition at line 380 of file block_sparsity_pattern.h. | https://www.dealii.org/developer/doxygen/deal.II/classBlockSparsityPatternBase.html | CC-MAIN-2019-47 | en | refinedweb |
kig
#include <bogus_imp.h>
Detailed Description
This ObjectImp is a BogusImp containing only a string value.
Definition at line 167 of file bogus_imp.h.
Member Typedef Documentation
Definition at line 176 of file bogus_imp.h.
Constructor & Destructor Documentation
Construct a new StringImp containing the string d.
Definition at line 54 of file bogus_imp.cc.
Member Function Documentation
Reimplemented from ObjectImp.
Definition at line 208 of file bogus_imp.cc.
Returns a copy of this ObjectImp.
The copy is an exact copy. Changes to the copy don't affect the original.
Reimplemented in TestResultImp.
Definition at line 69 of file bogus_imp.cc.
Get hold of the contained data.
Definition at line 186 of file bogus_imp.h.
Returns true if this ObjectImp is equal to rhs.
This function checks whether rhs is of the same ObjectImp type, and whether it contains the same data as this ObjectImp.
It is used e.g. by the KigCommand stuff to see what the user has changed during a move.
Reimplemented in TestResultImp.
Definition at line 175 of file bogus_imp.cc.
Reimplemented from ObjectImp.
Definition at line 103 of file bogus_imp.cc.
Set the contained data.
Definition at line 190 of file bogus_imp.h.
Returns the ObjectImpType representing the StringImp type.
Definition at line 220 of file bogus_imp.cc.
Returns the lowermost ObjectImpType that this object is an instantiation of.
E.g. if you want to get a string containing the internal name of the type of an object, you can do:
Reimplemented in TestResultImp.
Definition at line 255 of file bogus_imp.cc.
Reimplemented in TestResultImp.
Definition at line 133 of file bogus_imp.cc.
The documentation for this class was generated from the following files:
Documentation copyright © 1996-2020 The KDE developers.
Generated on Fri Jan 17 2020 03:27:12 by doxygen 1.8.7 written by Dimitri van Heesch, © 1997-2006
KDE's Doxygen guidelines are available online. | https://api.kde.org/4.14-api/kdeedu-apidocs/kig/html/classStringImp.html | CC-MAIN-2020-05 | en | refinedweb |
KmPlot
#include <function.h>
Detailed Description
This is the non-visual mathematical expression.
- Note
- when adding new member variables, make sure to update operator != and operator =.
Definition at line 238 of file function.h.
Member Enumeration Documentation
Definition at line 241 of file function.h.
Constructor & Destructor Documentation
Definition at line 278 of file function.cpp.
Definition at line 294 of file function.cpp.
Member Function Documentation
The full function expression, e.g.
"f(x,k)=(x+k)(x-k)".
Definition at line 301 of file function.h.
- Returns
- true if the fstr looks like "f(x) = ..."
- false if the fstr looks like "y = ..." (note that this depends on the type of equation, so if this is a Cartesian equation and the fstr looks like "a = ..." (not y) then it'll be considered a function, even if it isn't a very useful one.
Definition at line 343 of file function.cpp.
- Returns
- the name of the function, e.g. for the cartesian function f(x)=x^2, this would return "f".
Definition at line 317 of file function.cpp.
Definition at line 549 of file function.cpp.
Assigns the value in
other to this equation.
Definition at line 556 of file function.cpp.
- Returns
- the order of the differential equations.
Definition at line 299 of file function.cpp.
- Returns
- the name of the parameter variable (or a blank string if a parameter is not used).
Definition at line 466 of file function.cpp.
Definition at line 277 of file function.h.
- Returns
- the number of plus-minus symbols in the equation.
Definition at line 311 of file function.cpp.
The current plus-minus signature (true for plus, false for minus).
Definition at line 340 of file function.h.
- Parameters
-
- Returns
- whether
fstrcould be parsed correctly. Note that if it was not parsed correctly, then this will return false and this class will not be updated.
Definition at line 476 of file function.cpp.
- See also
- pmSignature.
Definition at line 542 of file function.cpp.
The type of function.
Definition at line 256 of file function.h.
Updates m_variables.
Definition at line 375 of file function.cpp.
- Returns
- whether the function accepts a parameter in addition to the x (and possibly y) variables.
Definition at line 292 of file function.h.
- Returns
- a list of variables, e.g. {x} for "f(x)=y", and {x,y,k} for "f(x,y,k)=(x+k)(y+k)".
Definition at line 287 of file function.h.
Member Data Documentation
For differential equations, all the states.
Definition at line 335 of file function.h.
Definition at line 354 of file function.h.
Definition at line 355 of file function.h.
Definition at line 356 of file function.h.
Definition at line 353 of file function.h.
Definition at line 352 of file function.h.
Cached list of variables.
Updated when setFstr is called.
Definition at line 360 of file function.h.
Pointer to the allocated memory for the tokens.
Definition at line 269 of file function.h.
Array index to the token.
Definition at line 273 of file function.h.
The documentation for this class was generated from the following files:
Documentation copyright © 1996-2020 The KDE developers.
Generated on Fri Jan 17 2020 03:39:06 by doxygen 1.8.7 written by Dimitri van Heesch, © 1997-2006
KDE's Doxygen guidelines are available online. | https://api.kde.org/4.x-api/kdeedu-apidocs/kmplot/kmplot/html/classEquation.html | CC-MAIN-2020-05 | en | refinedweb |
Background: Noda Time and C# 8
Note: this blog post was written based on experimentation with Visual Studio 2019 preview 2.2. It’s possible that some of the details here will change over time.
C# 8 is nearly here. At least, it’s close enough to being “here” that there are preview builds of Visual Studio 2019 available that support it. Unsurprisingly, I’ve been playing with it quite a bit.
In particular, I’ve been porting the Noda Time source code to use the new C# 8 features. The master branch of the repo is currently the code for Noda Time 3.0, which won’t be shipping (as a GA release) until after C# 8 and Visual Studio 2019 have fully shipped, so it’s a safe environment in which to experiment.
While it’s possible that I’ll use other C# 8 features in the future, the two C# 8 features that impact Noda Time most are nullable reference types and switch expressions. Both sets of changes are merged into master now, but the pull requests are still available so you can see just the changes:
The switch expressions PR is much simpler than the nullable reference types one. It’s entirely an implementation detail… although admittedly one that confused docfx, requiring a few of those switch expressions to be backed out or moved in a later PR.
Nullable reference types are a much, much bigger deal. They affect the public API, so they need to be treated much more carefully, and the changes end up being spread far wide throughout the codebase. That’s why the switch expression PR is a single commit, whereas nullable reference types is split into 14 commits – mostly broken up by project.
Reviewing the public API of a nullable reference type change
So I’m now in a situation where I’ve got nullable reference type support in Noda Time. Anyone consuming the 3.0 build (and there’s an alpha available for experimentation purposes) from C# 8 will benefit from the extra information that can now be expressed about parameters and return values. Great!
But how can I be confident in the changes to the API? My process for making the change in the first place was to enable nullable reference types and see what warnings were created. That’s a great starting point, but it doesn’t necessarily catch everything. In particular, although I started with the main project (the one that creates NodaTime.dll), I found that I needed to make more changes later on, as I modified other projects.
Just because your code compiles without any warnings with nullable reference types enabled doesn’t mean it’s “correct” in terms of the API you want to expose.
For example, consider this method:
public static string Identity(string input) => input;
That’s entirely valid C# 7 code, and doesn’t require any changes to compile, warning-free, in C# 8 with nullable reference types enabled. But it may not be what you actually want to expose. I’d argue that it should look like one of these:
// Allowing null input, and producing nullable output public static string? Identity(string? input) => input; // Preventing null input, and producing non-nullable output public static string Identity(string input) { // Convenience method for nullity checking. Preconditions.CheckNotNull(input, nameof(input)); return input; }
If you were completely diligent when writing tests for the code before C# 8, it should be obvious which is required – because you’d presumably have something like:
[Test] public void Identity_AcceptsNull() { Assert.IsNull(Identity(null)); }
That test would have produced a warning in C# 8, and would have suggested that the null-permissive API is the one you wanted. But maybe you forgot to write that test. Maybe the test you would have written was one that would have shown up a need to put that precondition in. It’s entirely possible that you write much more comprehensive tests than I do, but I suspect most of us have some code that isn’t explicitly tested in terms of its null handling.
The important part take-away here is that even code that hasn’t changed in appearance can change meaning in C# 8… so you really need to review any public APIs. How do you do that? Well, you could review the entire public API surface you’re exposing, of course. For many libraries that would be the simplest approach to take, as a “belt and braces” attitude to review. For Noda Time that’s less appropriate, as so much of the API only deals in value types. While a full API review would no doubt be useful in itself, I just don’t have the time to do it right now.
Instead, what I want to review is any API element which is impacted by the C# 8 change – even if the code itself hasn’t changed. Fortunately, that’s relatively easy to do.
Enter NullableAttribute
The C# 8 compiler applies a new attribute to every API element which is affected by nullability. As an example of what I mean by this, consider the following code which uses the
#nullable directive to control the nullable context of the code.
public class Test { #nullable enable public void X(string input) {} public void Y(string? input) {} #nullable restore #nullable disable public void Z(string input) {} #nullable restore }
The C# 8 compiler creates an internal
NullableAttribute class within the assembly (which I assume it wouldn’t if we were targeting a framework that already includes such an attribute) and applies the attribute anywhere it’s relevant. So the above code compiles to the same IL as this:
using System.Runtime.CompilerServices; public class Test { public void X([Nullable((byte) 1)] string input) {} public void Y([Nullable((byte) 2)] string input) {} public void Z(string input) {}} }
Note how the parameter for
Z doesn’t have the attribute at all, because that code is still oblivious to nullable reference types. But both
X and
Y have the attribute applied to their parameters – just with different arguments to describe the nullability. 1 is used for not-null; 2 is used for nullable.
That makes it relatively easy to write a tool to display every part of a library’s API that relates to nullable reference types – just find all the members that refer to
NullableAttribute, and filter down to public and protected members. It’s slightly annoying that
NullableAttribute doesn’t have any properties; code to analyze an assembly needs to find the appropriate
CustomAttributeData and examine the constructor arguments. It’s awkward, but not insurmountable.
I’ve started doing exactly that in the Noda Time repository, and got it to the state where it’s fine for Noda Time’s API review. It’s a bit quick and dirty at the moment. It doesn’t show protected members, or setter-only properties, or handle arrays, and there are probably other things I’ve forgotten about. I intend to improve the code over time and probably move it to my Demo Code repository at some point, but I didn’t want to wait until then to write about
NullableAttribute.
But hey, I’m all done, right? I’ve just explained how
NullableAttribute works, so what’s left? Well, it’s not quite as simple as I’ve shown so far.
NullableAttribute in more complex scenarios
It would be oh-so-simple if each parameter or return type could just be nullable or non-nullable. But life gets more complicated than that, with both generics and arrays. Consider a method called
GetNames() returning a list of strings. All of these are valid:
// Return value is non-null, and elements aren't null List<string> GetNames() // Return value is non-null, but elements may be null List<string?> GetNames() // Return value may be null, but elements aren't null List<string>? GetNames() // Return value may be null, and elements may be null List<string?>? GetNames()
So how are those represented in IL? Well,
NullableAttribute has one constructor accepting a single
byte for simple situations, but another one accepting
byte[] for more complex ones like this. Of course,
List<string> is still relatively simple – it’s just a single top-level generic type with a single type argument. For a more complex example, imagine
Dictionary<List<string?>, string[]?> . (A non-nullable reference to a dictionary where each key is a not-null list of nullable strings, and each value is a possibly-null array of non-nullable elements. Ouch.)
The layout of
NullableAttribute in these cases can be thought of in terms of a pre-order traversal of a tree representing the type, where generic type arguments and array element types are leaves in the tree. The above example could be thought of as this tree:
Dictionary<,> (not null) / \ / \ List<> (not null) Array (nullable) | | | | string (nullable) string (not null)
The pre-order traversal of that tree gives us these values:
- Not null (dictionary)
- Not null (list)
- Nullable (string)
- Nullable (array)
- Not null (string)
So a parameter declared with that type would be decorated like this:
[Nullable(new byte[] { 1, 1, 2, 2, 1 })]
But wait, there’s more!
NullableAttribute in simultaneously-complex-and-simple scenarios
The compiler has one more trick up its sleeve. When all the elements in the tree are “not null” or all elements in the tree are “nullable”, it simply uses the constructor with the single-byte parameter instead. So
Dictionary<List<string>, string[]> would be decorated with
Nullable[(byte) 1] and
Dictionary<List<string?>?, string?[]?>? would be decorated with
Nullable[(byte) 2].
(Admittedly,
Dictionary<,> doesn’t permit null keys anyway, but that’s an implementation detail.)
Conclusion
The C# 8 feature of nullable reference types is a really complicated one. I don’t think we’ve seen anything like this since async/await. This post has just touched on one interesting implementation detail. I’m sure there’ll be more posts on nullability over the next few months… | https://codeblog.jonskeet.uk/2019/02/ | CC-MAIN-2020-05 | en | refinedweb |
hcreate()
Create a hash search table
Synopsis:
#include <search.h> int hcreate( size_t nel );
Arguments:
- nel
- An estimate of the maximum number of entries that the table will contain. The algorithm might adjust this number upward in order to obtain certain mathematically favorable circumstances.
Library:
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
Description::
0 if there isn't enough space available to allocate the table. | https://developer.blackberry.com/playbook/native/reference/com.qnx.doc.neutrino.lib_ref/topic/h/hcreate.html | CC-MAIN-2020-05 | en | refinedweb |
What are field-programmable gate arrays (FPGA) and how to deploy
APPLIES TO:
Basic edition
Enterprise edition (Upgrade to Enterprise edition)
This article provides an introduction to field-programmable gate arrays (FPGA), and shows you how to deploy your models using Azure Machine Learning.
FPGAs on Azure inference (or model scoring) requests. Asynchronous requests (batching) aren't needed. Batching can cause latency, because more data needs to be processed.. Using this FPGA-enabled hardware architecture, trained neural networks run quickly and with lower latency. Azure can parallelize pre-trained deep neural networks (DNN) across FPGAs to scale out your service. The DNNs can be pre-trained, as a deep featurizer for transfer learning, or fine-tuned with updated weights.
FPGAs on Azure supports:
- Image classification and recognition scenarios
- TensorFlow deployment
- Intel FPGA hardware
These DNN models are currently available:
- ResNet 50
- ResNet 152
- DenseNet-121
- VGG-16
- SSD-VGG
FPGAs are available in these Azure regions:
- East US
- Southeast Asia
- West Europe
- West US 2
Important
To optimize latency and throughput, your client sending data to the FPGA model should be in one of the regions above (the one you deployed the model to).
The PBS Family of Azure VMs contains Intel Arria 10 FPGAs. It will show as "Standard PBS Family vCPUs" when you check your Azure quota allocation. The PB6 VM has six vCPUs and one FPGA, and it will automatically be provisioned by Azure ML as part of deploying a model to an FPGA. It is only used with Azure ML, and it cannot run arbitrary bitstreams. For example, you will not be able to flash the FPGA with bitstreams to do encryption, encoding, etc.
Scenarios and applications
Azure FPGAs are integrated with Azure Machine Learning. Microsoft uses FPGAs for DNN evaluation, Bing search ranking, and software defined networking (SDN) acceleration to reduce latency, while freeing CPUs for other tasks.
The following scenarios use FPGAs:
Example: Deploy models on FPGAs
You can deploy a model as a web service on FPGAs with Azure Machine Learning Hardware Accelerated Models. Using FPGAs provides ultra-low latency inference, even with a single batch size. Inference, or model scoring, is the phase where the deployed model is used for prediction, most commonly on production data.
Prerequisites
An Azure subscription. If you do not have one, create a free account before you begin. Try the free or paid version of Azure Machine Learning today.
FPGA quota. Use the Azure CLI to check whether you have quota:
az vm list-usage --location "eastus" -o table --query "[?localName=='Standard PBS Family vCPUs']"
Tip
The other possible locations are
southeastasia,
westeurope, and
westus2.
The command returns text similar to the following:
CurrentValue Limit LocalName -------------- ------- ------------------------- 0 6 Standard PBS Family vCPUs
Make sure you have at least 6 vCPUs under CurrentValue.
If you do not have quota, then submit a request at.
An Azure Machine Learning workspace and the Azure Machine Learning SDK for Python installed. For more information, see Create a workspace.
The Python SDK for hardware-accelerated models:
pip install --upgrade azureml-accel-models
1. Create and containerize models
This document will describe how to create a TensorFlow graph to preprocess the input image, make it a featurizer using ResNet 50 on an FPGA, and then run the features through a classifier trained on the ImageNet data set.
Follow the instructions to:
- Define the TensorFlow model
- Convert the model
- Deploy the model
- Consume the deployed model
- Delete deployed services to run on the FPGA.
Load Azure Machine Learning workspace
Load your Azure Machine Learning workspace.
import os import tensorflow as tf from azureml.core import Workspace ws = Workspace.from_config() print(ws.name, ws.resource_group, ws.location, ws.subscription_id, sep='\n')
Preprocess image
The input to the web service is a JPEG image. The first step is to decode the JPEG image and preprocess it. The JPEG images are treated as strings and the result are tensors that will be the input to the ResNet 50 model.
# Input images as a two-dimensional tensor containing an arbitrary number of images represented a strings import azureml.accel.models.utils as utils tf.reset_default_graph() in_images = tf.placeholder(tf.string) image_tensors = utils.preprocess_array(in_images) print(image_tensors.shape)
Load featurizer
Initialize the model and download a TensorFlow checkpoint of the quantized version of ResNet50 to be used as a featurizer. You may replace "QuantizedResnet50" in the code snippet below with by importing other deep neural networks:
- QuantizedResnet152
- QuantizedVgg16
- Densenet121
from azureml.accel.models import QuantizedResnet50 save_path = os.path.expanduser('~/models') model_graph = QuantizedResnet50(save_path, is_frozen=True) feature_tensor = model_graph.import_graph_def(image_tensors) print(model_graph.version) print(feature_tensor.name) print(feature_tensor.shape)
Add classifier
This classifier has been trained on the ImageNet data set. Examples for transfer learning and training your customized weights are available in the set of sample notebooks.
classifier_output = model_graph.get_default_classifier(feature_tensor) print(classifier_output)
Save the model
Now that the preprocessor, ResNet 50 featurizer, and the classifier have been loaded, save the graph and associated variables as a model.
model_name = "resnet50" model_save_path = os.path.join(save_path, model_name) print("Saving model in {}".format(model_save_path)) with tf.Session() as sess: model_graph.restore_weights(sess) tf.saved_model.simple_save(sess, model_save_path, inputs={'images': in_images}, outputs={'output_alias': classifier_output})
Save input and output tensors
The input and output tensors that were created during the preprocessing and classifier steps will be needed for model conversion and inference.
input_tensors = in_images.name output_tensors = classifier_output.name print(input_tensors) print(output_tensors)
Important
Save the input and output tensors because you will need them for model conversion and inference requests.
The available models and the corresponding default classifier output tensors are below, which is what you would use for inference if you used the default classifier.
- Resnet50, QuantizedResnet50
output_tensors = "classifier_1/resnet_v1_50/predictions/Softmax:0"
- Resnet152, QuantizedResnet152
output_tensors = "classifier/resnet_v1_152/predictions/Softmax:0"
- Densenet121, QuantizedDensenet121
output_tensors = "classifier/densenet121/predictions/Softmax:0"
- Vgg16, QuantizedVgg16
output_tensors = "classifier/vgg_16/fc8/squeezed:0"
- SsdVgg, QuantizedSsdVgg
output_tensors = ['ssd_300_vgg/block4_box/Reshape_1:0', 'ssd_300_vgg/block7_box/Reshape_1:0', 'ssd_300_vgg/block8_box/Reshape_1:0', 'ssd_300_vgg/block9_box/Reshape_1:0', 'ssd_300_vgg/block10_box/Reshape_1:0', 'ssd_300_vgg/block11_box/Reshape_1:0', 'ssd_300_vgg/block4_box/Reshape:0', 'ssd_300_vgg/block7_box/Reshape:0', 'ssd_300_vgg/block8_box/Reshape:0', 'ssd_300_vgg/block9_box/Reshape:0', 'ssd_300_vgg/block10_box/Reshape:0', 'ssd_300_vgg/block11_box/Reshape:0']
Register model
Register the model by using the SDK with the ZIP file in Azure Blob storage. Adding tags and other metadata about the model helps you keep track of your trained models.
from azureml.core.model import Model registered_model = Model.register(workspace=ws, model_path=model_save_path, model_name=model_name) print("Successfully registered: ", registered_model.name, registered_model.description, registered_model.version, sep='\t')
If you've already registered a model and want to load it, you may retrieve it.
from azureml.core.model import Model model_name = "resnet50" # By default, the latest version is retrieved. You can specify the version, i.e. version=1 registered_model = Model(ws, name="resnet50") print(registered_model.name, registered_model.description, registered_model.version, sep='\t')
Convert model
Convert the TensorFlow graph to the Open Neural Network Exchange format (ONNX). You will need to provide the names of the input and output tensors, and these names will be used by your client when you consume the web service.
from azureml.accel import AccelOnnxConverter convert_request = AccelOnnxConverter.convert_tf_model( ws, registered_model, input_tensors, output_tensors) # If it fails, you can run wait_for_completion again with show_output=True. convert_request.wait_for_completion(show_output=False) # If the above call succeeded, get the converted model converted_model = convert_request.result print("\nSuccessfully converted: ", converted_model.name, converted_model.url, converted_model.version, converted_model.id, converted_model.created_time, '\n')
Create Docker image
The converted model and all dependencies are added to a Docker image. This Docker image can then be deployed and instantiated. Supported deployment targets include AKS in the cloud or an edge device such as Azure Data Box Edge. You can also add tags and descriptions for your registered Docker image.
from azureml.core.image import Image from azureml.accel import AccelContainerImage image_config = AccelContainerImage.image_configuration() # Image name must be lowercase image_name = "{}-image".format(model_name) image = Image.create(name=image_name, models=[converted_model], image_config=image_config, workspace=ws) image.wait_for_creation(show_output=False)
List the images by tag and get the detailed logs for any debugging.
for i in Image.list(workspace=ws): print('{}(v.{} [{}]) stored at {} with build log {}'.format( i.name, i.version, i.creation_state, i.image_location, i.image_build_log_uri))
2. Deploy to cloud or edge
Deploy to the cloud
To deploy your model as a high-scale production web service, use Azure Kubernetes Service (AKS). You can create a new one using the Azure Machine Learning SDK, CLI, or Azure Machine Learning studio.
from azureml.core.compute import AksCompute, ComputeTarget # Specify the Standard_PB6s Azure VM and location. Values for location may be "eastus", "southeastasia", "westeurope", or "westus2”. If no value is specified, the default is "eastus". prov_config = AksCompute.provisioning_configuration(vm_size = "Standard_PB6s", agent_count = 1, location = "eastus") aks_name = 'my-aks-cluster' # Create the cluster aks_target = ComputeTarget.create(workspace=ws, name=aks_name, provisioning_configuration=prov_config)
The AKS deployment may take around 15 minutes. Check to see if the deployment succeeded.
aks_target.wait_for_completion(show_output=True) print(aks_target.provisioning_state) print(aks_target.provisioning_errors)
Deploy the container to the AKS cluster.
from azureml.core.webservice import Webservice, AksWebservice # For this deployment, set the web service configuration without enabling auto-scaling or authentication for testing aks_config = AksWebservice.deploy_configuration(autoscale_enabled=False, num_replicas=1, auth_enabled=False) aks_service_name = 'my-aks-service' aks_service = Webservice.deploy_from_image(workspace=ws, name=aks_service_name, image=image, deployment_config=aks_config, deployment_target=aks_target) aks_service.wait_for_deployment(show_output=True)
Test the cloud service
The Docker image supports gRPC and the TensorFlow Serving "predict" API. Use the sample client to call into the Docker image to get predictions from the model. Sample client code is available:
If you want to use TensorFlow Serving, you can download a sample client.
# Using the grpc client in Azure ML Accelerated Models SDK package from azureml.accel import PredictionClient address = aks_service.scoring_uri ssl_enabled = address.startswith("https") address = address[address.find('/')+2:].strip('/') port = 443 if ssl_enabled else 80 # Initialize Azure ML Accelerated Models client client = PredictionClient(address=address, port=port, use_ssl=ssl_enabled, service_name=aks_service.name)
Since this classifier was trained on the ImageNet data set, map the classes to human-readable labels.
import requests classes_entries = requests.get( "").text.splitlines() # Score image with input and output tensor names results = client.score_file(path="./snowleopardgaze.jpg", input_name=input_tensors, outputs=output_tensors) # map results [class_id] => [confidence] results = enumerate(results) # sort results by confidence sorted_results = sorted(results, key=lambda x: x[1], reverse=True) # print top 5 results for top in sorted_results[:5]: print(classes_entries[top[0]], 'confidence:', top[1])
Clean-up the service
Delete your web service, image, and model (must be done in this order since there are dependencies).
aks_service.delete() aks_target.delete() image.delete() registered_model.delete() converted_model.delete()
Deploy to a local edge server
All Azure Data Box Edge devices contain an FPGA for running the model. Only one model can be running on the FPGA at one time. To run a different model, just deploy a new container. Instructions and sample code can be found in this Azure Sample.
Secure FPGA web services
To secure your FPGA web services, see the Secure web services document.
Next steps
Feedback | https://docs.microsoft.com/en-us/azure/machine-learning/how-to-deploy-fpga-web-service | CC-MAIN-2020-05 | en | refinedweb |
Related
How To Use the Bottle Micro Framework to Develop Python Web Apps
Python is an excellent language for web programming due to its flexibility and high-level functionality. Web frameworks can make programming web applications much simpler because they connect many of the components necessary for a robust web interface.
While some web frameworks attempt to provide everything that a user might want to use to develop an application, others try to stay out of the way while taking care of the important, difficult to implement issues. Bottle is a Python framework that falls into the second category. It is extremely lightweight, but makes it very easy to develop applications quickly.
In this guide, we will cover how to set up and use Bottle to create simple web applications on an Ubuntu 12.04 server.
How To Install Bottle
Python, the programming language that Bottle is built for, comes installed on Ubuntu by default.
Install and Activate a Virtual Environment
We will install the
virtualenv package to isolate our Python project from the system’s Python environment.
We can install this easily from the repositories:
sudo apt-get update sudo apt-get install python-virtualenv
The virtualenv software allows us to create a separate, contained environment for our Python projects that will not affect the entire OS. We are going to create a projects folder in our home directory and then create a virtual environment within this folder:
mkdir ~/projects cd ~/projects virtualenv --no-site-packages venv
This creates a directory called
venv within the projects directory. It has installed some Python utilities within this folder and created a directory structure to install additional tools.
We must activate the virtual environment before beginning to work on our project:
source venv/bin/activate
The command prompt will change to reflect the fact that we are operating in a virtual environment now. If you need to exit the virtual environment, you can type this at any time:
deactivate
Do not deactivate your virtual environment at this point. just installing the Bottle package:
pip install bottle
After the process completes, we should have the ability to use the Bottle framework within our applications. editor, create a Python application called
hello.py:
nano hello.py
Within this file, we are going to first import some functionality from the Bottle package. This will allow us to use the framework tools within our application:
from bottle import route, run
This line tells our program that we want to import the route and run modules from the Bottle package.
The run module that we are importing can be used to run the application in a development server, which is great for quickly seeing the results of your program.
The route module patter
/hello:
from bottle import route, run @route('/hello')
This route decorator matches the URL
/hello when that path is requested on the server. The function that directly follows will be executed when this matches: parameter specifies the port that this will be using.
Save and close the file.
We can run this application by typing this:
python hello.py
You can visit this application in your web browser by going to your IP address, followed by the port we chose to run on (8080), followed by the route we created (/hello):
You can stop the server at any time by typing “CTRL-C” in the terminal window.
Implement the MVC Design Paradigm
We have now implemented our first application. It was certainly simple, but it doesn’t really implement MVC principles, or do anything particularly interesting. Let’s try to make a more complicated application this time.
Create the Model
Let’s that our application can implement.
Install the SQLite software in Ubuntu to make sure we have the software available to create and interact with these databases:
sudo apt-get install sqlite
We also need to download and install the Bottle plugin that will allow us to use these databases:
pip install bottle-sqlite
Now that we have the components, we can create a simple database that we can store our data in. We will create a Python file that will generate a SQLite database with some data in it when we run the script. We could do this in the Python interpreter, but this way makes it easy to repeat.
nano picnic_data.py
Here, we import the SQLite package. Then, we can execute a command that creates our table and inserts data in our table. Finally, we commit the changes:()
Save and close the file.
We can execute the file, which will create a database file called
picnic.db within our current directory:
python picnic_data.py
Our model portion of our program is now fairly complete. We can see that our model will dictate a little bit how our control portion must function to interact with our data.
Create the Controller
Now that we have a database created, we can start to develop our main application. This will mainly implement our controller functionality. It will also be the file that most closely resembles our first application.
Create a file called
picnic.py to store our main application:
nano picnic.py. Finally, it returns to formatted output to our user.
import sqlite3 from bottle import route, run, template @route('/picnic') def show_picnic(): db = sqlite3.connect('picnic.db') c = db.cursor() c.execute("SELECT item,quant FROM picnic") data = c.fetchall() c.close() output = template('bring_to_picnic', rows=data) return output
Finally, we need to put connect to the databasetopicnic.tpl to format the data. It passes the “data” variable as the template variable “rows”.
We will create this template file in the next section.
Create the:
nano bring_to_picnic.tpl
In this file, we can mix HTML and programming. Ours will be very simple. It will use a loop to create a table, which we will populate with our model data:
<h1>Things to bring to our picnic</h1> <table> <tr><th>Item</th><th>Quantity</th></tr> %for row in rows: <tr> %for col in row: <td>{{col}}</td> %end </tr> %end </table>
This will render our page in HTML. The templating language that we see here is basically Python. The “rows” variable that we passed to the template is available to use when designing the output.
We can type lines of Python by preceding them with “%”. We can access variables within the HTML by using the “{{var}}” syntax.
Save and close the file.
Viewing the Results
Our application is now complete and we can start the program by calling Python on the main file:
python picnic.py
We can see the results by visiting our IP address and port followed by the URL route we created:
. One easy way to find plugins is by using the
pip search bottle command. This will give you an idea of some of the more popular options.
10 Comments | https://www.digitalocean.com/community/tutorials/how-to-use-the-bottle-micro-framework-to-develop-python-web-apps | CC-MAIN-2020-05 | en | refinedweb |
Table of Contents or in a database (§7.2). Packages that are stored in a file system may using qualified names. This can prevent the conflicts that would otherwise occur if two development groups happened to pick the same package name and these packages were later to be used in a single program.
The members of a package are its subpackages and all the top level class types (§7.6, §8 (Classes)) and top level interface types (§9 (Interfaces)) declared in all the compilation units (§7.3) of the package.
For example, in the Java SE platform API:
The package
java has
subpackages
awt,
applet,
io,
lang,
net,
and
util, but no compilation units.
The package
java.awt has a
subpackage named
image, as well as a number
of compilation units containing declarations of class and
interface types.
If the fully qualified name
(§6.7) of a package is
P,
and
Q is a subpackage of
P,
then
P.Q is the fully qualified name of the
subpackage, and furthermore denotes a
package.
A package may not contain two members of the same name, or a compile-time error results.
Here are some examples:
Because the package
java.awt
has a subpackage
image, it cannot (and does
not) contain a declaration of a class or interface type
named
image.
If there is a package
named
mouse and a member
type
Button in that package (which then might
be referred to as
mouse.Button), then there
cannot be any package with the fully qualified
name
mouse.Button
or
mouse.Button.Click.
If
com.nighthacks.java.jag is
the fully qualified name of a type, then there cannot be any
package whose fully qualified name is
either
com.nighthacks.java.jag
or
com.nighthacks.java.jag.scrabble.
It is however possible for members of different packages to have the same simple name. For example, it is possible to declare a package:
package vector; public class Vector { Object[] vec; }
that has as a member a
public class
named
Vector, even though the package
java.util
also declares a class named
Vector. These two class
types are different, reflected by the fact that they have different
fully qualified names (§6.7). The fully qualified
name of this example
Vector
is
vector.Vector,
whereas
java.util.Vector is the fully qualified
name of the
Vector class included in the
Java SE platform. Because the package
vector contains a
class named
Vector, it cannot also have a
subpackage named
Vector.
The hierarchical naming structure for packages is intended to be convenient for organizing related packages in a conventional manner, but has no significance in itself other than the prohibition against a package having a subpackage with the same simple name as a top level type (§7.6) declared in that package.
For example, there is no special access relationship
between a package named
oliver and another package
named
oliver.twist, or between packages
named
evelyn.wood
and
evelyn.waugh. That is, the code in a package
named
oliver.twist has no better access to the
types declared within package
oliver than code in
any other package.
Each host system determines how packages and compilation units are created and stored.
Each host system also determines which compilation units are observable (§7.3) in a particular compilation. The observability of compilation units in turn determines which packages are observable, and which packages are in scope.
In simple implementations of the Java SE platform, packages and compilation units may be stored in a local file system. Other implementations may store them using a distributed file system or some form of database.
If a host system stores packages and compilation units in a database, then the database must not impose the optional restrictions (§7.6) on compilation units permissible.
As an extremely simple example of storing packages in a file system, all the packages and source and binary code in a project might be stored in a single directory and its subdirectories. Each immediate subdirectory of this directory would represent a top level package, that is, one whose fully qualified name consists of a single simple name. Each further level of subdirectory would represent a subpackage of the package represented by the containing directory, and so on.
The directory might contain the following immediate subdirectories:
com gls jag java wnj
where directory
java would
contain the Java SE platform packages; the
directories
jag,
gls,
and
wnj might contain packages that three of the
authors of this specification created for their personal use and to
share with each other within this small group; and the
directory
com would contain packages procured from
companies that used the conventions described in
§6.1 to generate unique names for their
packages.
Continuing the example, the
directory
java would contain, among others, the
following subdirectories:
applet awt io lang net util
corresponding to the
packages
java.applet,
java.awt,
java.io,
java.lang,
java.net, and
java.util that are defined as part
of the Java SE platform API.
Still continuing the example, if we were to look
inside the directory
util, we might see the
following files:
BitSet.java Observable.java BitSet.class Observable.class Date.java Observer.java Date.class Observer.class ...
where each of the
.java files
contains the source for a compilation unit (§7.3)
that contains the definition of a class or interface whose binary
compiled form is contained in the
corresponding
.class file.
Under this simple organization of packages, an implementation of the Java SE platform would transform a package name into a pathname by concatenating the components of the package name, placing a file name separator (directory indicator) between adjacent components.
For example, if this simple organization were used
on an operating system where the file name separator
is
/, the package name:
jag.scrabble.board
would be transformed into the directory name:
jag/scrabble/board
@
character followed by four hexadecimal digits giving the numeric value
of the character, as in
the
\uxxxx escape
(§3.3).
Under this convention, the package name:
children.activities.crafts.papierM\u00e2ch\u00e9
which can also be written using full Unicode as:
children.activities.crafts.papierMâché
might be mapped to the directory name:
children/activities/crafts/papierM@00e2ch@00e9
If the
@ character is not a valid
character in a file name for some given host file system, then some
other character that is not valid in a identifier could be used
instead.
CompilationUnit is the goal symbol (§2.1) for the syntactic grammar (§2.3) of Java programs. It is defined by the following productions:
A compilation unit consists of three parts, each of which is optional:
A
package declaration
(§7.4), giving the fully qualified name
(§6.7) of the package to which the
compilation unit belongs.
A compilation unit that
has no
package declaration is part of an unnamed package
(§7.4.2).
import declarations
(§7.5) that allow types from other packages
and
static members of types to be referred to using their simple
names.
Top level type declarations (§7.6) of class and interface types.
Every
compilation unit implicitly imports every
public type name declared
in the predefined package
java.lang, as if the
declaration
import java.lang.*; appeared at the
beginning of each compilation unit immediately after any
package
statement. As a result, the names of all those types are available as
simple names in every compilation unit.
All the compilation units of
the predefined package
java and
its subpackages
lang
and
io are always
observable.
For all other packages, the host system determines which compilation units are observable.
The observability of a compilation unit influences the observability of its package (§7.4.3).
Types declared in different compilation units can depend on each other, circularly. A Java compiler must arrange to compile all such types at the same time.
A
package declaration appears
within a compilation unit to indicate the package to which the
compilation unit belongs.
A package declaration in a compilation unit specifies the name (§6.2) of the package to which the compilation unit belongs.
packageIdentifier {
.Identifier}
;
The
package name mentioned in a
package declaration must be the fully
qualified name of the package (§6.7).
The scope and shadowing of a package declaration is specified in §6.3 and §6.4.
The rules for annotation modifiers on a package declaration are specified in §9.7.4 and §9.7.5.
At most
one annotated
package declaration is permitted for a given
package.
The manner in which this restriction is enforced
must, of necessity, vary from implementation to implementation. The
following scheme is strongly recommended for file-system-based
implementations: The sole annotated
package declaration, if it
exists, is placed in a source file
called
package-info.java in the directory
containing the source files for the package. This file does not
contain the source for a class called
package-info.java; indeed it would be illegal for
it to do so, as
package-info is not a legal
identifier. Typically
package-info.java contains
only a
package declaration, preceded immediately by the annotations
on the package. While the file could technically contain the source
code for one or more classes with package access, it would be very bad
form.
It is recommended
that
package-info.java, if it is present, take the
place of
package.html
for
javadoc and other similar documentation
generation systems. If this file is present, the documentation
generation tool should look for the package documentation comment
immediately preceding the (possibly annotated)
package declaration
in
package-info.java. In this
way,
package-info.java becomes the sole repository
for package-level annotations and documentation. If, in future, it
becomes desirable to add any other package-level information, this
file should prove a convenient home for this information.
A
compilation unit that has no
package declaration is part of
an unnamed package.
Unnamed packages are provided by the Java SE platform principally for convenience when developing small or temporary applications or when just beginning development.
An unnamed package cannot
have subpackages, since the syntax of a
package declaration always
includes a reference to a named top level package.
An implementation of the Java SE platform must support at least one unnamed package. An implementation may support more than one unnamed package, but is not required to do so. Which compilation units are in each unnamed package is determined by the host system.
Example 7.4.2-1.
The compilation unit:
class FirstCall { public static void main(String[] args) { System.out.println("Mr. Watson, come here. " + "I want you."); } }
defines a very simple compilation unit as part of an unnamed package.
In implementations of the Java SE.
A package is observable if and only if either:
A compilation unit containing a declaration of the package is observable (§7.3).
A subpackage of the package is observable.
The
packages
java,
java.lang, and
java.io are always
observable.
One can conclude this from the rule above and from
the rules of observable compilation units, as follows. The predefined
package
java.lang declares the class
Object, so the compilation
unit for
Object is always observable
(§7.3). Hence, the
java.lang package is
observable (§7.4.3), and
the
java package also. Furthermore, since
Object
is observable, the array type
Object
[] implicitly
exists. Its superinterface
java.io.Serializable (§10.1)
also exists, hence the
java.io package is observable.
An import declaration allows
a named type or a
static member to be referred to by a simple name
(§6.2) that consists of a single
identifier.
Without the use of an
appropriate import declaration, the only way to refer to a type
declared in another package, or a
static member of another type, is
to use a fully qualified name (§6.7).
A single-type-import declaration (§7.5.1) imports a single named type, by mentioning its canonical name (§6.7).
A type-import-on-demand declaration (§7.5.2) imports all the accessible types (§6.6) of a named type or named package as needed, by mentioning the canonical name of a type or package.
A single-static-import
declaration (§7.5.3) imports all accessible
static members with a given name from a type, by giving its
canonical name.
A static-import-on-demand
declaration (§7.5.4) imports all accessible
static members of a named type as needed, by mentioning the
canonical name of a type.
The scope and shadowing of a type or member imported by these declarations is specified in §6.3 and §6.4.
An
import declaration makes types or members
available by their simple names only within the compilation unit that
actually contains the
import declaration. The scope of the type(s)
or member(s) introduced by an
import declaration specifically does
not include other compilation units in the same package, other
import declarations in the current compilation unit, or a
package
declaration in the current compilation unit (except for the
annotations of a
package declaration).
A single-type-import declaration imports a single type by giving its canonical name, making it available under a simple name in the class and interface declarations of the compilation unit in which the single-type-import declaration appears.
The TypeName must be the canonical name of a class type, interface type, enum type, or annotation type (§6.7).
The name must be qualified (§6.5.5.2), or a compile-time error occurs.
It is a compile-time error if the named type is not accessible (§6.6).
If two single-type-import declarations in the same compilation unit attempt to import types with the same simple name, then a compile-time error occurs, unless the two types are the same type, in which case the duplicate declaration is ignored.
If the
type imported by the single-type-import declaration is declared in the
compilation unit that contains the
import declaration, the
import
declaration is ignored.
If a single-type-import declaration imports a type whose simple name is n, and the compilation unit also declares a top level type (§7.6) whose simple name is n, a compile-time error occurs.
If a compilation unit contains both a single-type-import declaration that imports a type whose simple name is n, and a single-static-import declaration (§7.5.3) that imports a type whose simple name is n, a compile-time error occurs.
Example 7.5.1-1. Single-Type-Import
import java.util.Vector;
causes the simple name
Vector to
be available within the class and interface declarations in a
compilation unit. Thus, the simple name
Vector
refers to the type declaration
Vector in the
package
java.util in all places where it is not shadowed
(§6.4.1) or obscured
(§6.4.2) by a declaration of a field, parameter,
local variable, or nested type declaration with the same name.
Note that the actual declaration
of
java.util.Vector is generic
(§8.1.2). Once imported, the
name
Vector can be used without qualification in a
parameterized type such as
Vector<String>, or
as the raw type
Vector. A related limitation of the
import declaration is that a nested type declared inside a generic
type declaration can be imported, but its outer type is always
erased.
Example 7.5.1-2. Duplicate Type Declarations
This program:
import java.util.Vector; class Vector { Object[] vec; }
causes a compile-time error because of the duplicate
declaration of
Vector, as does:
import java.util.Vector; import myVector.Vector;
where
myVector is a package
containing the compilation unit:
package myVector; public class Vector { Object[] vec; }
Example 7.5.1-3. No Import of a Subpackage
Note that an
import statement cannot import a
subpackage, only a type.
For example, it does not work to try to import
java.util and then use the name
util.Random to
refer to the type
java.util.Random:
import java.util; class Test { util.Random generator; } // incorrect: compile-time error
Example 7.5.1-4. Importing a Type Name that is also a Package Name
Package names and type names are usually different
under the naming conventions described in
§6.1. Nevertheless, in a contrived example where
there is an unconventionally-named package
Vector,
which declares a public class whose name
is
Mosquito:
package Vector; public class Mosquito { int capacity; }
and then the compilation unit:
package strange; import java.util.Vector; import Vector.Mosquito; class Test { public static void main(String[] args) { System.out.println(new Vector().getClass()); System.out.println(new Mosquito().getClass()); } }
the single-type-import declaration importing
class
Vector from package
java.util does not
prevent the package name
Vector from appearing and
being correctly recognized in subsequent
import declarations. The
example compiles and produces the output:
class java.util.Vector class Vector.Mosquito
A type-import-on-demand declaration allows all accessible types of a named package or type to be imported as needed.
importPackageOrTypeName
.
*
;
The PackageOrTypeName must be the canonical name (§6.7) of a package, a class type, an interface type, an enum type, or an annotation type.
If the PackageOrTypeName denotes a type (§6.5.4), then the name must be qualified (§6.5.5.2), or a compile-time error occurs.
It is a compile-time error if the named package or type is not accessible (§6.6).
It is not
a compile-time error to name either
java.lang or the named package of
the current compilation unit in a type-import-on-demand
declaration. The type-import-on-demand declaration is ignored in such
cases.
Two or more type-import-on-demand declarations in the same compilation unit may name the same type or package. All but one of these declarations are considered redundant; the effect is as if that type was imported only once.
If a
compilation unit contains both a type-import-on-demand declaration and
a static-import-on-demand declaration (§7.5.4)
that name the same type, the effect is as if the
static member types
of that type (§8.5, §9.5)
were imported only once.
Example 7.5.2-1. Type-Import-on-Demand
import java.util.*;
causes the simple names of all
public types
declared in the package
java.util to be available within the class
and interface declarations of the compilation unit. Thus, the simple
name
Vector refers to the
type
Vector in the package
java.util in all places
in the compilation unit where that type declaration is not shadowed
(§6.4.1) or obscured
(§6.4.2).
The declaration might be shadowed by a
single-type-import declaration of a type whose simple name
is
Vector; by a type
named
Vector and declared in the package to which
the compilation unit belongs; or any nested classes or
interfaces.
The declaration might be obscured by a declaration
of a field, parameter, or local variable
named
Vector.
(It would be unusual for any of these conditions to occur.)
A single-static-import
declaration imports all accessible
static members with a
given simple name from a type. This makes these
static members
available under their simple name in the class and interface
declarations of the compilation unit in which the single-static-import
declaration appears.
import
staticTypeName
.Identifier
;).
The
Identifier must name at least one
static member of the named
type. It is a compile-time error if there is
no
static member of that name, or if all of the
named members are not accessible.
It is permissible for one single-static-import declaration to import several fields or types with the same name, or several methods with the same name and signature.
If a single-static-import declaration imports a type whose simple name is n, and the compilation unit also declares a top level type (§7.6) whose simple name is n, a compile-time error occurs.
If a compilation unit contains both a single-static-import declaration that imports a type whose simple name is n, and a single-type-import declaration (§7.5.1) that imports a type whose simple name is n, a compile-time error occurs.
A static-import-on-demand
declaration allows all accessible
static
members of a named type to be imported as
needed.).
Two or more static-import-on-demand declarations in the same compilation unit may name the same type; the effect is as if there was exactly one such declaration.
Two or more static-import-on-demand declarations in the same compilation unit may name the same member; the effect is as if the member was imported exactly once.
It is permissible for one static-import-on-demand declaration to import several fields or types with the same name, or several methods with the same name and signature.
If a
compilation unit contains both a static-import-on-demand declaration
and a type-import-on-demand declaration (§7.5.2)
that name the same type, the effect is as if the
static member types
of that type (§8.5, §9.5)
were imported only once.
A top level type declaration declares a top level class type (§8 (Classes)) or a top level interface type (§9 (Interfaces))..
In the absence of an access modifier, a
top level type has package access: it is accessible only within
compilation units of the package in which it is declared
(§6.6.1). A type may be declared
public to
grant access to the type from code in other packages.
It is a
compile-time error if a top level type declaration contains any one of
the following access modifiers:
protected,
private, or
static.
It is a compile-time error if the name of a top level type appears as the name of any other top level class or interface type declared in the same package.
The scope and shadowing of a top level type is specified in §6.3 and §6.4.
The fully qualified name of a top level type is specified in §6.7.
Example 7.6-1. Conflicting Top Level Type Declarations
package test; import java.util.Vector; class Point { int x, y; } interface Point { // compile-time error #1 int getR(); int getTheta(); } class Vector { Point[] pts; } // compile-time error #2
Here, the first compile-time error is caused by the
duplicate declaration of the name
Point as both a
class and an interface in the same package. A second compile-time
error is the attempt to declare the name
Vector both by a class type declaration and by a
single-type-import declaration.
Note, however, that it is not an error for the name of a class to also name a type that otherwise might be imported by a type-import-on-demand declaration (§7.5.2) in the compilation unit (§7.3) containing the class declaration. Thus, in this program:
package test; import java.util.*; class Vector {} // not a compile-time error).
Example 7.6-2. Scope of Top Level Types
package points; class Point { int x, y; // coordinates PointColor color; // color of this point Point next; // next point with this color static int nPoints; } class PointColor { Point first; // first point with this color PointColor(int color) { this.color = color; } private int color; // color components }
This program defines two classes that use each other
in the declarations of their class members. Because the class
types
Point and
PointColor have
all the type declarations in package
points,
including all those in the current compilation unit, as their scope,
this program compiles correctly. That is, forward reference is not a
problem.
Example 7.6-3. Fully Qualified Names
class Point { int x, y; }
In this code, the class
Point is
declared in a compilation unit with no
package statement, and
thus
Point is its fully qualified name, whereas in
the code:
package vista; class Point { int x, y; }
the fully qualified name of the
class
Point is
vista.Point. (The
package name
vista is suitable for local or
personal use; if the package were intended to be widely distributed,
it would be better to give it a unique package name
(§6.1).)
An implementation of the Java SE):
import java.util.Vector;
then within that compilation unit, the simple
name
Vector and the fully qualified
name
java.util.Vector refer to the same
type.
If and only
if packages are stored in a file system (§7. | https://docs.oracle.com/javase/specs/jls/se8/html/jls-7.html | CC-MAIN-2020-05 | en | refinedweb |
Sh here in Jordan and try to stay here illegally,” he says. “We already have our own unemployment problems, especially among the younger generation.”
Al-Faisal’s group has come up with a technological solution, utilizing drones and voice recognition. “The voice is the key. Native Arabic speakers in West Bank and East Bank populations have noticeable differences in their use of voiceless sibilant fricatives.” He points to a quadcopter drone, its four propellers whirring as it hovers in place. “We have installed a high-quality audio system onto these drones, with a speaker and directional microphones. The drone is autonomous. It presents a verbal challenge to the suspect. When the suspect answers, digital signal processing computes the coefficient of palatalization. East Bank natives have a mean palatalization coefficient of 0.85. With West Bank natives it is only 0.59. If the coefficient is below the acceptable threshold, the drone immobilizes the suspect with nonlethal methods until the nearest patrol arrives.”
One of Al-Faisal’s students presses a button and the drone zooms up and off to the west, embarking on another pilot run. If Al-Faisal’s system, the SIBFRIC-2000, proves to be successful in these test runs, it is likely to be used on a larger scale — so that limited Jordanian resources can cover a wider area to patrol for illegal immigrants. “Two weeks ago we caught a group of eighteen illegals with SIBFRIC. Border Security would never have found them. It’s not like in America, where you can discriminate against people because they look different. East Bank, West Bank — we all look the same. We sound different. I am confident the program will work.”
But some residents of the region are skeptical. “My brother was hit by a drone stunner on one of these tests,” says a 25-year old bus driver, who did not want his name to be used for this story. “They said his voice was wrong. Our family has lived in Jordan for generations. We are proud citizens! How can you trust a machine?”
Al-Faisal declines to give out statistics on how many cases like this have occurred, citing national security concerns. “The problem is being overstated. Very few of these incidents have occurred, and in each case the suspect is unharmed. It does not take long for Border Security to verify citizenship once they arrive.”
Others say there are rumors of voice coaches in Nablus and Ramallah, helping desperate refugees to beat the system. Al-Faisal is undeterred. “Ha ha ha, this is nonsense. SIBFRIC has a perfect ear; it can hear the slightest nuances of speech. You cannot cheat the computer.”
When asked how many drones are in the pilot program, Al-Faisal demures. “More than five,” he says, “fewer than five hundred.” Al-Faisal hopes to ramp up the drone program starting in 2021.
Shame Old Shtory, Shame Old Shong and Dansh
Does this story sound familiar? It might. Here’s a similar one from several thousand years earlier: (King James Version)
This is the reputed biblical origin of the term shibboleth, which has come to mean any distinguishing cultural behavior that can be used to identify a particular group, not just a linguistic one. The Ephraimites couldn’t say SHHHH and it became a dead giveaway of their origin.
In this article, we’re going to talk about the shibboleth and several other diagnostic tests which have one of two outcomes — pass or fail, yes or no, positive or negative — and some of the implications of using such a test. And yes, this does impact embedded systems. We’re going to spend quite a bit of time looking at a specific example in mathematical terms. If you don’t like math, just skip it whenever it gets too “mathy”.
A quantitative shibboleth detector
So let’s say we did want to develop a technological test for separating out people who couldn’t say SHHH, and we had some miracle algorithm to evaluate a “palatalization coefficient” \( P_{SH} \) for voiceless sibilant fricatives. How would we pick the threshold?
Well, the first thing to do is try to model the system somehow. We took a similar approach in an earlier article on design margin, where our friend the Oracle computed probability distributions for boiler pressures. Let’s do the same here. Suppose the Palestinians (which we’ll call Group 1) have a \( P_{SH} \) which roughly follows a Gaussian distribution with mean μ = 0.59 and standard deviation σ = 0.06, and the Jordanians (which we’ll call Group 2) have a \( P_{SH} \) which is also a Gaussian distribution with a mean μ = 0.85 and a standard deviation σ = 0.03.
import numpy as np import matplotlib.pyplot as plt %matplotlib inline import scipy.stats from collections import namedtuple class Gaussian1D(namedtuple('Gaussian1D','mu sigma name id color')): """ Normal (Gaussian) distribution of one variable """ def pdf(self, x): """ probabiliy density function """ return scipy.stats.norm.pdf(x, self.mu, self.sigma) def cdf(self, x): """ cumulative distribution function """ return scipy.stats.norm.cdf(x, self.mu, self.sigma) d1 = Gaussian1D(0.59, 0.06, 'Group 1 (Palestinian)','P','red') d2 = Gaussian1D(0.85, 0.03, 'Group 2 (Jordanian)','J','blue') def show_binary_pdf(d1, d2, x, fig=None, xlabel=None): if fig is None: fig = plt.figure(figsize=(6,3)) ax=fig.add_subplot(1,1,1) for d in (d1,d2): ax.plot(x,d.pdf(x),label=d.name,color=d.color) ax.legend(loc='best',labelspacing=0,fontsize=11) if xlabel is not None: ax.set_xlabel(xlabel, fontsize=15) ax.set_ylabel('probability density') return ax show_binary_pdf(d1,d2,x=np.arange(0,1,0.001), xlabel='$P_{SH}$');
Hmm, we have a dilemma here. Both groups have a small possibility that the \( P_{SH} \) measurement is in the 0.75-0.8 range, and this makes it harder to distinguish. Suppose we decided to set the threshold at 0.75. What would be the probability of our conclusion being wrong?
for d in (d1,d2): print '%-25s %f' % (d.name, d.cdf(0.75))
Group 1 (Palestinian) 0.996170 Group 2 (Jordanian) 0.000429
Here the
cdf method calls the
scipy.stats.norm.cdf function to compute the cumulative distribution function, which is the probability that a given sample from the distribution will be less than a given amount. So there’s a 99.617% chance that Group 1’s \( P_{SH} < 0.75 \), and a 0.0429% chance that Group 2’s \( P_{SH} < 0.75 \). One out of every 261 samples from Group 1 will pass the test (though we were hoping for them to fail) — this is known as a false negative, because a condition that exists (\( P_{SH} < 0.75 \)) remains undetected. One out of every 2331 samples from Group 2 will fail the test (though we were hoping for them to pass) — this is known as a false positive, because a condition that does not exist (\( P_{SH} < 0.75 \)) is mistakenly detected.
The probabilities of false positive and false negative are dependent on the threshold:
for threshold in [0.72,0.73,0.74,0.75,0.76,0.77,0.78]: c = [d.cdf(threshold) for d in (d1,d2)] print("threshold=%.2f, Group 1 false negative=%7.5f%%, Group 2 false positive=%7.5f%%" % (threshold, 1-c[0], c[1])) threshold = np.arange(0.7,0.801,0.005) false_positive = d2.cdf(threshold) false_negative = 1-d1.cdf(threshold) figure = plt.figure() ax = figure.add_subplot(1,1,1) ax.semilogy(threshold, false_positive, label='false positive', color='red') ax.semilogy(threshold, false_negative, label='false negative', color='blue') ax.set_ylabel('Probability') ax.set_xlabel('Threshold $P_{SH}$') ax.legend(labelspacing=0,loc='lower right') ax.grid(True) ax.set_xlim(0.7,0.8);
threshold=0.72, Group 1 false negative=0.01513%, Group 2 false positive=0.00001% threshold=0.73, Group 1 false negative=0.00982%, Group 2 false positive=0.00003% threshold=0.74, Group 1 false negative=0.00621%, Group 2 false positive=0.00012% threshold=0.75, Group 1 false negative=0.00383%, Group 2 false positive=0.00043% threshold=0.76, Group 1 false negative=0.00230%, Group 2 false positive=0.00135% threshold=0.77, Group 1 false negative=0.00135%, Group 2 false positive=0.00383% threshold=0.78, Group 1 false negative=0.00077%, Group 2 false positive=0.00982%
If we want a lower probability of false positives (fewer Jordanians detained for failing the test) we can do so by lowering the threshold, but at the expense of raising the probability of false negatives (more Palestinians unexpectedly passing the test and not detained), and vice-versa.
Factors used in choosing a threshold
There is a whole science around binary-outcome tests, primarily in the medical industry, involving sensivity and specificity, and it’s not just a matter of probability distributions. There are two other aspects that make a huge difference in determining a good test threshold:
- base rate — the probability of the condition actually being true, sometimes referred as prevalence in medical diagnosis
- the consequences of false positives and false negatives
Both of these are important because they affect our interpretation of false positive and false negative probabilities.
Base rate
The probabilities we calculated above are conditional probabilities — in our example, we calculated the probability that a person known to be from the Palestinian population passed the SIBFRIC test, and the probability that a person known to be from the Jordanian population failed the SIBFRIC tests.
It’s also important to consider the joint probability distribution — suppose that we are trying to detect a very uncommon condition. In this case the false positive rate will be amplified relative to the false negative rate. Let’s say we have some condition C that has a base rate of 0.001, or one in a thousand, and there is a test with a false positive rate of 0.2% and a false negative rate of 5%. This sounds like a really bad test: we should balance the probabilities by lowering the false negative rate and allowing a higher false positive rate. The net incidence of false positives for C will be 0.999 × 0.002 = 0.001998, and the net incidence of false negatives will be 0.001 × 0.05 = 0.00005. If we had one million people we test for condition C:
- 1000 actually have condition C
- 950 people are correctly diagnosed has having C
- 50 people will remain undetected (false negatives)
- 999000 do not actually have condition C
- 997002 people are correctly diagnosed as not having C
- 1998 people are incorrectly diagnosed as having C (false positives)
The net false positive rate is much higher than the net false negative rate, and if we had a different test with a false positive rate of 0.1% and a false negative rate of 8%, this might actually be better, even though the conditional probabilities of false positives and false negatives look even more lopsided. This is known as the false positive paradox.
Consequences
Let’s continue with our hypothetical condition C with a base rate of 0.001, and the test that has a false positive rate of 0.2% and a false negative rate of 5%. And suppose that the consequences of false positives are unnecessary hospitalization and the consequences of false negatives are certain death:
- 997002 diagnosed as not having C → relief
- 1998 incorrectly diagnosed as having C → unnecessary hospitalization, financial cost, annoyance
- 950 correctly diagnosed as having C → treatment, relief
- 50 incorrectly diagnosed as not having C → death
If the test can be changed, we might want to reduce the false negative, even if it raises the net false positive rate higher. Would lowering 50 deaths per million to 10 deaths per million be worth it if it raises the false positive rate of unnecessary hospitalization from 1998 per million to, say, 5000 per million? 20000 per million?
Consequences can rarely be compared directly; more often we have an apples-to-oranges comparison like death vs. unnecessary hospitalization, or allowing criminals to be free vs. incarcerating the innocent. If we want to handle a tradeof quantitatively, we’d need to assign some kind of metric for the consequences, like assigning a value of \$10 million for an unnecessary death vs. \$10,000 for an unnecessary hospitalization — in such a case we can minimize the net expected loss over an entire population. Otherwise it becomes an ethical question. In jurisprudence there is the idea of Blackstone’s ratio: “It is better that ten guilty persons escape than that one innocent suffer.” But the post-2001 political climate in the United States seems to be that detaining the innocent is more desirable than allowing terrorists or illegal immigrants to remain at large. Mathematics alone won’t help us out of these quandaries.
Optimizing a threshold
OK, so suppose we have a situation that can be described completely in mathematical terms:
- \( A \): Population A can be measured with parameter \( x \) with probability density function \( f_A(x) \) and cumulative density function \( F_A(x) = \int\limits_{-\infty}^x f_A(u) \, du \)
- \( B \): Population B can be measured with parameter \( x \) with PDF \( f_B(x) \) and CDF \( F_B(x) = \int\limits_{-\infty}^x f_B(u) \, du \)
- All samples are either in \( A \) or \( B \):
- \( A \) and \( B \) are disjoint (\( A \cap B = \varnothing \))
- \( A \) and \( B \) are collectively exhaustive (\( A \cup B = \Omega \), where \( \Omega \) is the full sample space, so that \( P(A \ {\rm or}\ B) = 1 \))
- Some threshold \( x_0 \) is determined
- For any given sample \( s \) that is in either A or B (\( s \in A \) or \( s \in B \), respectively), parameter \( x_s \) is compared with \( x_0 \) to determine an estimated classification \( a \) or \( b \):
- \( a \): if \( x_s > x_0 \) then \( s \) is likely to be in population A
- \( b \): if \( x_s \le x_0 \) then \( s \) is likely to be in population B
- Probability of \( s \in A \) is \( p_A \)
- Probability of \( s \in B \) is \( p_B = 1-p_A \)
- Value of various outcomes:
- \( v_{Aa} \): \( s \in A, x_s > x_0 \), correctly classified in A
- \( v_{Ab} \): \( s \in A, x_s \le x_0 \), incorrectly classified in B
- \( v_{Ba} \): \( s \in B, x_s > x_0 \), incorrectly classified in A
- \( v_{Bb} \): \( s \in B, x_s \le x_0 \), correctly classified in B
The expected value over all outcomes is
$$\begin{aligned} E[v] &= v_{Aa}P(Aa)+v_{Ab}P(Ab) + v_{Ba}P(Ba) + v_{Bb}P(Bb)\cr &= v_{Aa}p_A P(a\ |\ A) \cr &+ v_{Ab}p_A P(b\ |\ A) \cr &+ v_{Ba}p_B P(a\ |\ B) \cr &+ v_{Bb}p_B P(b\ |\ B) \end{aligned}$$
These conditional probabilities \( P(a\ |\ A) \) (denoting the probability of the classification \( a \) given that the sample is in \( A \)) can be determined with the CDF functions; for example, if the sample is in A then \( P(a\ |\ A) = P(x > x_0\ |\ A) = 1 - P(x \le x_0) = 1 - F_A(x_0) \), and once we know that, then we have
$$\begin{aligned} E[v] &= v_{Aa}p_A (1-F_A(x_0)) \cr &+ v_{Ab}p_A F_A(x_0) \cr &+ v_{Ba}p_B (1-F_B(x_0)) \cr &+ v_{Bb}p_B F_B(x_0) \cr E[v] &= p_A \left(v_{Ab} + (v_{Aa}-v_{Ab})\left(1-F_A(x_0)\right)\right) \cr &+ p_B \left(v_{Bb} + (v_{Ba}-v_{Bb})\left(1-F_B(x_0)\right)\right) \end{aligned}$$
\( E[v] \) is actually a function of the threshold \( x_0 \), and we can locate its maximum value by determining points where its partial derivative \( {\partial E[v] \over \partial {x_0}} = 0: \)
$$0 = {\partial E[v] \over \partial {x_0}} = p_A(v_{Ab}-v_{Aa})f_A(x_0) + p_B(v_{Bb}-v_{Ba})f_B(x_0)$$
where the \( \rho \) are ratios for probability and for value tradeoffs:
$$\begin{aligned} \rho_p &= p_B/p_A \cr \rho_v &= -\frac{v_{Bb}-v_{Ba}}{v_{Ab}-v_{Aa}} \end{aligned}$$
One interesting thing about this equation is that since probabilities \( p_A \) and \( p_B \) and PDFs \( f_A \) and \( f_B \) are positive, this means that \( v_{Bb}-v_{Ba} \) and \( v_{Ab}-v_{Aa} \) must be opposite signs, otherwise… well, let’s see:
Case study: Embolary Pulmonism
Suppose we have an obscure medical condition; let’s call it an embolary pulmonism, or EP for short. This must be treated within 48 hours, or the patient can transition in minutes from seeming perfectly normal, to a condition in which a small portion of their lungs degrade rapidly, dissolve, and clog up blood vessels elsewhere in the body, leading to extreme discomfort and an almost certain death. Before this rapid decline, the only symptoms are a sore throat and achy eyes.
We’re developing an inexpensive diagnostic test \( T_1 \) (let’s suppose it costs \$1) where the patient looks into a machine and it takes a picture of the patient’s eyeballs and uses machine vision to come up with some metric \( x \) that can vary from 0 to 100. We need to pick a threshold \( x_0 \) such that if \( x > x_0 \) we diagnose the patient with EP.
Let’s consider some math that’s not quite realistic:
- condition A: patient has EP
- condition B: patient does not have EP
- incidence of EP in patients complaining of sore throats and achy eyes: \( p_A = \) 0.004% (40 per million)
- value of Aa (correct diagnosis of EP): \( v_{Aa}=- \) \$100000
- value of Ab (false negative, patient has EP, diagnosis of no EP): \( v_{Ab}=0 \)
- value of Ba (false positive, patient does not have EP, diagnosis of EP): \( v_{Ba}=- \)\$5000
- value of Bb (correct diagnosis of no EP): \( v_{Bb} = 0 \)
So we have \( v_{Ab}-v_{Aa} = \) \$100000 and \( v_{Bb}-v_{Ba} = \) \$5000, for a \( \rho_v = -0.05, \) which implies that we’re looking for a threshold \( x_0 \) where \( \frac{f_A(x_0)}{f_B(x_0)} = -0.05\frac{p_B}{p_A} \) is negative, and that never occurs with any real probability distributions. In fact, if we look carefully at the values \( v \), we’ll see that when we diagnose a patient with EP, it always has a higher cost: If we correctly diagnose them with EP, it costs \$100,000 to treat. If we incorrectly diagnose them with EP, it costs \$5,000, perhaps because we can run a lung biopsy and some other fancy test to determine that it’s not EP. Whereas if we give them a negative diagnosis, it doesn’t cost anything. This implies that we should always prefer to give patients a negative diagnosis. So we don’t even need to test them!
Patient: “Hi, Doc, I have achy eyes and a sore throat, do I have EP?”
Doctor: (looks at patient’s elbows studiously for a few seconds) “Nope!”
Patient: (relieved) “Okay, thanks!”
What’s wrong with this picture? Well, all the values look reasonable except for two things. First, we haven’t included the \$1 cost of the eyeball test… but that will affect all four outcomes, so let’s just state that the values \( v \) are in addition to the cost of the test. The more important issue is the false negative, the Ab case, where the patient is diagnosed incorrectly as not having EP, and it’s likely the patient will die. Perhaps the hospital’s insurance company has estimated a cost of \$10 million per case to cover wrongful death civil suits, in which case we should be using \( v _ {Ab} = -\$10^7 \). So here’s our revised description:
- condition A: patient has EP
- condition B: patient does not have EP
- incidence of EP in patients complaining of sore throats and achy eyes: \( p_A = \) 0.004% (40 per million)
- value of Aa (correct diagnosis of EP): \( v_{Aa}=- \) \$100000 (rationale: mean cost of treatment)
- value of Ab (false negative, patient has EP, diagnosis of no EP): \( v_{Ab}=-\$10^7 \) (rationale: mean cost of resulting liability due to high risk of death)
- value of Ba (false positive, patient does not have EP, diagnosis of EP): \( v_{Ba}=- \)\$5000 (rationale: mean cost of additional tests to confirm)
- value of Bb (correct diagnosis of no EP): \( v_{Bb} = 0 \) (rationale: no further treatment needed)
The equation for choosing \( x_0 \) then becomes
$$\begin{aligned} \rho_p &= p_B/p_A = 0.99996 / 0.00004 = 24999 \cr \rho_v &= -\frac{v_{Bb}-v_{Ba}}{v_{Ab}-v_{Aa}} = -5000 / {-9.9}\times 10^6 \approx 0.00050505 \cr {f_A(x_0) \over f_B(x_0)} &= \rho_p\rho_v \approx 12.6258. \end{aligned}$$
Now we need to know more about our test. Suppose that the results of the eye machine test have normal probability distributions with \( \mu=55, \sigma=5.3 \) for patients with EP and \( \mu=40, \sigma=5.9 \) for patients without EP.
x = np.arange(0,100,0.1) dpos = Gaussian1D(55, 5.3, 'A (EP-positive)','A','red') dneg = Gaussian1D(40, 5.9, 'B (EP-negative)','B','green') show_binary_pdf(dpos, dneg, x, xlabel = 'test result $x$');
Yuck. This doesn’t look like a very good test; there’s a lot of overlap between the probability distributions.
At any rate, suppose we pick a threshold \( x_0=47 \); what kind of false positive / false negative rates will we get, and what’s the expected overall value?
import IPython.core.display from IPython.display import display, HTML def analyze_binary(dneg, dpos, threshold): """ Returns confusion matrix """ pneg = dneg.cdf(threshold) ppos = dpos.cdf(threshold) return np.array([[pneg, 1-pneg], [ppos, 1-ppos]]) def show_binary_matrix(confusion_matrix, threshold, distributions, outcome_ids, ppos, vmatrix, special_format=None): if special_format is None: special_format = {} def cellinfo(c, p, v): # joint probability = c*p jp = c*p return (c, jp, jp*v) def rowcalc(i,confusion_row): """ write this for rows containing N elements, not just 2 """ p = ppos if (i == 1) else (1-ppos) return [cellinfo(c,p,vmatrix[i][j]) for j,c in enumerate(confusion_row)] Jfmtlist = special_format.get('J') cfmtlist = special_format.get('c') vfmtlist = special_format.get('v') try: if isinstance(vfmtlist, basestring): vfmt_general = vfmtlist else: vfmt_general = vfmtlist[0] except: vfmt_general = '%.3f' def rowfmt(row,dist): def get_format(fmt, icell, default): if fmt is None: return default if isinstance(fmt,basestring): return fmt return fmt[icell] or default def cellfmt(icell): Jfmt = get_format(Jfmtlist, icell, '%.7f') cfmt = get_format(cfmtlist, icell, '%.5f') vfmt = get_format(vfmtlist, icell, '%.3f') return '<td>'+cfmt+'<br>J='+Jfmt+'<br>wv='+vfmt+'</td>' return '<th>'+dist.name+'</th>' + ''.join( (cellfmt(i) % cell) for i,cell in enumerate(row)) rows = [rowcalc(i,row) for i,row in enumerate(confusion_matrix)] vtot = sum(v for row in rows for c,J,v in row) if not isinstance (threshold, basestring): threshold = 'x_0 = %s' % threshold return HTML(('<p>Report for threshold \\(%s \\rightarrow E[v]=\\)' +vfmt_general+'</p>')%(threshold,vtot) +'<table><tr><td></td>'+ ''.join('<th>%s</th>' % id for id in outcome_ids) +'</tr>' +''.join('<tr>%s</tr>' % rowfmt(row,dist) for row,dist in zip(rows,distributions)) +'</table>') threshold = 47 C = analyze_binary(dneg, dpos, threshold) show_binary_matrix(C, threshold, [dneg, dpos], 'ba', 40e-6,[[0,-5000],[-1e7, -1e5]], special_format={'v':'$%.2f'})
Report for threshold \(x_0 = 47 \rightarrow E[v]=\)$-618.57
Here we’ve shown a modified confusion matrix showing for each of the four outcomes the following quantities:
- Conditional probability of each outcome: \( \begin{bmatrix}P(b\ |\ B) & P(a\ |\ B) \cr P(b\ |\ A) & P(a\ |\ A)\end{bmatrix} \) — read each entry like \( P(a\ |\ B) \) as “the probability of \( a \), given that \( B \) is true”, so that the numbers in each row add up to 1
- J: Joint probability of the outcome: \( \begin{bmatrix}P(Bb) & P(Ba) \cr P(Ab) & P(Aa)\end{bmatrix} \) — read each entry like \( P(Ba) \) as “The probability that \( B \) and \( a \) are true”, so that the numbers in the entire matrix add up to 1
- wv: Weighted contribution to expected value = joint probability of the outcome × its value
For example, if the patient does not have EP, there’s about an 88.2% chance that they will be diagnosed correctly and an 11.8% chance that the test will produce a false positive. If the patient does have EP, there’s about a 6.6% chance the test will produce a false negative, and a 93.4% chance the test will correctly diagnose that they have EP.
The really interesting thing here is the contribution to expected value. Remember, the false negative (Ab) is really bad, since it has a value of \$10 million, but it’s also very rare because of the low incidence of EP and the fact that condtional probability of a false negative is only 6.6%. Whereas the major contribution to expected value comes from the false positive case (Ba) that occurs in almost 11.8% of the population.
We should be able to use a higher threshold to reduce the expected cost over the population:
for threshold in [50, 55]: C = analyze_binary(dneg, dpos, threshold) display(show_binary_matrix(C, threshold, [dneg, dpos], 'ba', 40e-6,[[0,-5000],[-1e7, -1e5]], special_format={'v':'$%.2f'}))
Report for threshold \(x_0 = 50 \rightarrow E[v]=\)$-297.62
Report for threshold \(x_0 = 55 \rightarrow E[v]=\)$-229.52
The optimal threshold should probably be somewhere between \( x_0=50 \) and \( x_0=55 \), since in one case the contributions to expected value is mostly from the false positive case, and in the other from the false negative case. If we have a good threshold, these contributions are around the same order of magnitude from the false positive and false negative cases. (They won’t necessarily be equal, though.)
To compute this threshold we are looking for
$$\rho = {f_A(x_0) \over f_B(x_0)} = \rho_p\rho_v \approx 12.6258.$$
We can either solve it using numerical methods, or try to solve analytically using the normal distribution probability density
$$f(x) = \frac{1}{\sigma\sqrt{2\pi}}e^{-(x-\mu)^2/2\sigma^2}$$
which gives us
$$\frac{1}{\sigma_A}e^{-(x_0-\mu_A)^2/2\sigma_A{}^2} = \frac{1}{\sigma_B}\rho e^{-(x_0-\mu_B)^2/2\sigma_B{}^2}$$
and taking logs, we get
$$-\ln\sigma_A-(x_0-\mu_A)^2/2\sigma_A{}^2 = \ln\rho -\ln\sigma_B-(x_0-\mu_B)^2/2\sigma_B{}^2$$
If we set \( u = x_0 - \mu_A \) and \( \Delta = \mu_B - \mu_A \) then we get
$$-u^2/2\sigma_A{}^2 = \ln\rho\frac{\sigma_A}{\sigma_B} -(u-\Delta)^2/2\sigma_B{}^2$$
$$-\sigma_B{}^2u^2 = 2\sigma_A{}^2\sigma_B{}^2\left(\ln\rho\frac{\sigma_A}{\sigma_B} \right) -\sigma_A{}^2(u^2 - 2\Delta u + \Delta^2)$$
which simplifies to \( Au^2 + Bu + C = 0 \) with
$$\begin{aligned} A &= \sigma_B{}^2 - \sigma_A{}^2 \cr B &= 2\Delta\sigma_A{}^2 \cr C &= 2\sigma_A{}^2\sigma_B{}^2\left(\ln\rho\frac{\sigma_A}{\sigma_B}\right) - \Delta^2\sigma_A{}^2 \end{aligned}$$
We can solve this with the alternate form of the quadratic formula \( u = \frac{2C}{-B \pm \sqrt{B^2-4AC}} \) which can compute the root(s) even with \( A=0\ (\sigma_A = \sigma_B = \sigma) \), where it simplifies to \( u=-C/B=\frac{-\sigma^2 \ln \rho}{\mu_B - \mu_A} + \frac{\mu_B - \mu_A}{2} \) or \( x_0 = \frac{\mu_B + \mu_A}{2} - \frac{\sigma^2 \ln \rho}{\mu_B - \mu_A} \).
def find_threshold(dneg, dpos, ppos, vmatrix): num = -(1-ppos)*(vmatrix[0][0]-vmatrix[0][1]) den = ppos*(vmatrix[1][0]-vmatrix[1][1]) rho = num/den A = dneg.sigma**2 - dpos.sigma**2 ofs = dpos.mu delta = dneg.mu - dpos.mu B = 2.0*delta*dpos.sigma**2 C = (2.0 * dneg.sigma**2 * dpos.sigma**2 * np.log(rho*dpos.sigma/dneg.sigma) - delta**2 * dpos.sigma**2) if (A == 0): roots = [ofs-C/B] else: D = B*B-4*A*C roots = [ofs + 2*C/(-B-np.sqrt(D)), ofs + 2*C/(-B+np.sqrt(D))] # Calculate expected value, so that if we have more than one root, # the caller can determine which is better pneg = 1-ppos results = [] for i,root in enumerate(roots): cneg = dneg.cdf(root) cpos = dpos.cdf(root) Ev = (cneg*pneg*vmatrix[0][0] +(1-cneg)*pneg*vmatrix[0][1] +cpos*ppos*vmatrix[1][0] +(1-cpos)*ppos*vmatrix[1][1]) results.append((root,Ev)) return results find_threshold(dneg, dpos, 40e-6, [[0,-5000],[-1e7, -1e5]])
[(182.23914143860179, -400.00000000000006), (53.162644275683974, -212.51747111423805)]
threshold =53.1626 for x0 in [threshold, threshold-0.1, threshold+0.1]: C = analyze_binary(dneg, dpos, x0) display(show_binary_matrix(C, x0, [dneg, dpos], 'ba', 40e-6,[[0,-5000],[-1e7, -1e5]], special_format={'v':'$%.2f'}))
Report for threshold \(x_0 = 53.1626 \rightarrow E[v]=\)$-212.52
Report for threshold \(x_0 = 53.0626 \rightarrow E[v]=\)$-212.58
Report for threshold \(x_0 = 53.2626 \rightarrow E[v]=\)$-212.58
So it looks like we’ve found the threshold \( x_0 = 53.0626 \) to maximize expected value of all possible outcomes at \$-212.52. The vast majority (98.72%) of people taking the test don’t incur any cost or trouble beyond that of the test itself.
Still, it’s somewhat unsatisfying to have such a high false negative rate: over 36% of patients who are EP-positive are undetected by our test, and are likely to die. To put this into perspective, consider a million patients who take this test. The expected number of them for each outcome are
- 12842 will be diagnosed with EP but are actually EP-negative (false positive) and require \$5000 in tests to confirm
- 15 will be EP-positive but will not be diagnosed with EP (false negative) and likely to die
- 25 will be EP-positive and correctly diagnosed and incur \$100,000 in treatment
- the rest are EP-negative and correctly diagnosed.
That doesn’t seem fair to those 15 people, just to help reduce the false positive rate.
We could try skewing the test by assigning a value of \$100 million, rather than \$10 million, for the false negative case, because it’s really really bad:
find_threshold(dneg, dpos, 40e-6, [[0,-5000],[-1e8, -1e5]])
[(187.25596232479654, -4000.0000000000005), (48.145823389489138, -813.91495238472135)]
threshold = 48.1458 for x0 in [threshold, threshold-0.1, threshold+0.1]: C = analyze_binary(dneg, dpos, x0) display(show_binary_matrix(C, x0, [dneg, dpos], 'ba', 40e-6,[[0,-5000],[-1e8, -1e5]], special_format={'v':'$%.2f'}))
Report for threshold \(x_0 = 48.1458 \rightarrow E[v]=\)$-813.91
Report for threshold \(x_0 = 48.0458 \rightarrow E[v]=\)$-814.23
Report for threshold \(x_0 = 48.2458 \rightarrow E[v]=\)$-814.23
There, by moving the threshold downward by about 5 points, we’ve reduced the false negative rate to just under 10%, and the false positive rate is just over 8%. The expected value has nearly quadrupled, though, to \$-813.91, mostly because the false positive rate has increased, but also because the cost of a false negative is much higher.
Now, for every million people who take the test, we would expect
- around 83700 will have a false positive
- 36 will be correctly diagnosed with EP
- 4 will incorrectly remain undiagnosed and likely die.
Somehow that doesn’t sound very satisfying either. Can we do any better?
Idiot lights
Let’s put aside our dilemma of choosing a diagnosis threshold, for a few minutes, and talk about idiot lights. This term generally refers to indicator lights on automotive instrument panels, and apparently showed up in print around 1960. A July 1961 article by Phil McCafferty in Popular Science, called Let’s Bring Back the Missing Gauges, states the issue succinctly:
Car makers have you and me figured out for idiots — to the tune of about eight million dollars a year. This is the estimated amount that auto makers save by selling us little blink-out bulbs — known to fervent nonbelievers as “idiot lights” — on five million new cars instead of fitting them out with more meaningful gauges.
Not everything about blink-outs is bad. They do give a conspicuous warning when something is seriously amiss. But they don’t tell enough, or tell it soon enough to be wholly reliable. That car buyers aren’t happy is attested by the fact that gauge makers gleefully sell some quarter of a million accessory instruments a year to people who insist on knowing what’s going on under their hoods.
He goes on to say:
There’s little consolation in being told your engine has overheated after it’s done it. With those wonderful old-time gauges, a climbing needle gave you warning before you got into trouble.
The basic blink-out principle is contrary to all rules of safety. Blink-outs do not “fail safe.” The system operates on the assurance that all is well when the lights are off. If a bulb burns out while you’re traveling, you’ve lost your warning system.
Sure, most indicators remain on momentarily during starting, which makes it possible to check them for burnouts. But this has been known to have problems.
Consider the Midwestern husband who carefully lectured his young wife on the importance of watching gauges, then traded in the old car for a new blink-out model. Anxious to try it out, the wife fired it up the minute it arrived without waiting for her husband to come home. Panicked by three glaring lights, she assumed the worst, threw open the hood, and was met by the smell of new engine paint burning. Without hesitation, she popped off the cap and filled the crank-case to the brim—with water.
The big blink-out gripe is that the lights fail to tell the degree of what is taking place. Most oil-pressure blink-outs turn off at about 10 to 15 pounds’ pressure. Yet this is not nearly enough to lubricate an engine at 70 m.p.h.
The generator blink-out, unlike an ammeter, tells only whether or not the generator is producing current, not how much. You can be heading for a dead battery if you are using more current than the generator is producing. A battery can also be ruined by pouring an excessive charge into it, and over-production can kill a generator. Yet in all cases the generator is working, so the light is off.
The lights hide the secrets that a sinking, climbing, or fluctuating needle can reveal. Because of this, blink-outs are one of the greatest things that ever happened to a shady used-car dealer.
McCafferty seems to have been a bit of a zealot on this topic, publishing a November 1955 article in Popular Science, I Like the Gauges Detroit Left Out, though it doesn’t mention the term “idiot light”. The earliest use of “idiot light” in common media seems to be in the January 1960 issue of Popular Science, covering some automotive accessories like this volt-ammeter kit:
Interestingly enough, on the previous page is an article (TACHOMETER: You Can Assemble One Yourself with This \$14.95 Kit) showing a schematic and installation of a tachometer circuit that gets its input from one of the distributor terminals, and uses a PNP transistor to send a charge pulse to a capacitor and ammeter everytime the car’s distributor sends a high-voltage pulse to the spark plugs:
The earliest use of “idiot light” I could find in any publication is from a 1959 monograph on Standard Cells from the National Bureau of Standards which states
It is realized that the use of good-bad indicators must be approached with caution. Good-bad lights, often disparagingly referred to as idiot lights, are frequently resented by the technician because of the obvious implication. Furthermore, skilled technicians may feel that a good-bad indication does not give them enough information to support their intuitions. However, when all factors are weighed, good-bad indicators appear to best fit the requirements for an indication means that may be interpreted quickly and accurately by a wide variety of personnel whether trained or untrained.
Regardless of the term’s history, we’re stuck with these lights in many cases, and they’re a compromise between two principles:
- idiot lights are an effective and inexpensive method of catching the operator’s attention to one of many possible conditions that can occur, but they hide information by reducing a continuous value to a true/false indication
- a numeric result, shown by a dial-and-needle gauge or a numeric display, can show more useful information than an idiot light, but it is more expensive, doesn’t draw attention as easily as an idiot light, and it requires the operator to interpret the numeric value
¿Por qué no los dos?
If we don’t have an ultra-cost-sensitive system, why not have both? Computerized screens are very common these days, and it’s relatively easy to display both a PASS/FAIL or YES/NO indicator — for drawing attention to a possible problem — and a value that allows the operator to interpret the data.
Since 2008, new cars sold in the United States have been required to have a tire pressure monitoring system (TPMS). As a driver, I both love it and hate it. The TPMS is great for dealing with slow leaks before they become a problem. I carry a small electric tire pump that plugs into the 12V socket in my car, so if the TPMS light comes on, I can pull over, check the tire pressure, and pump up the one that has a leak to its normal range. I’ve had slow leaks that have lasted several weeks before they start getting worse. Or sometimes they’re just low pressure because it’s November or December and the temperature has decreased. What I don’t like is that there’s no numerical gauge. If my TPMS light comes on, I have no way to distinguish a slight decrease in tire pressure (25psi vs. the normal 30psi) vs. a dangerously low pressure (15psi), unless I stop and measure all four tires with a pressure gauge. I have no way to tell how quickly the tire pressure is decreasing, so I can decide whether to keep driving home and deal with it later, or whether to stop at the nearest possible service station. It would be great if my car had an information screen where I could read the tire pressure readings and decide what to do based on having that information.
As far as medical diagnostic tests go, using the extra information from a raw test score can be a more difficult decision, especially in cases where the chances and costs of both false positives and false negatives are high. In the EP example we looked at earlier, we had a 0-to-100 test with a threshold of somewhere in the 55-65 range. Allowing a doctor to use their judgment when interpreting this kind of a test might be a good thing, especially when the doctor can try to make use of other information. But as a patient, how am I to know? If I’m getting tested for EP and I have a 40 reading, my doctor can be very confident that I don’t have EP, whereas with a 75 reading, it’s a no-brainer to start treatment right away. But those numbers near the threshold are tricky.
Triage (¿Por qué no los tres?)
In medicine, the term triage refers to a process of rapid prioritization or categorization in order to determine which patients should be served first. The idea is to try to make the most difference, given limited resources — so patients who are sick or injured, but not in any immediate danger, may have to wait.
As an engineer, my colleagues and I use triage as a way to categorize issues so that we can focus only on the few that are most important. A couple of times a year we’ll go over the unresolved issues in our issue tracking database, to figure out which we’ll address in the near term. One of the things I’ve noticed is that our issues fall into three types:
- Issues which are obviously low priority — these are ones that we can look at in a few seconds and agree, “Oh, yeah, we don’t like that behavior but it’s just a minor annoyance and isn’t going to cause any real trouble.”
- Issues which are obviously high priority — these are ones that we can also look at in a few seconds and agree that we need to address them soon.
- Issues with uncertainty — we look at these and kind of sit and stare for a while, or have arguments within the group, about whether they’re important or not.
The ones in the last category take a lot of time, and slow this process down immensely. I would much rather come to a 30-second consensus of L/H/U (“low priority”/”high priority”/”uncertain”) and get through the whole list, then come back and go through the U issues one by one at a later date.
Let’s go back to our EP case, and use the results of our \$1 eyeball-photography test \( T_1 \), but instead of dividing our diagnosis into two outcomes, let’s divide it into three outcomes:
- Patients are diagnosed as EP-positive, with high confidence
- Patients for which the EP test is “ambivalent” and it is not possible to distinguish between EP-positive and EP-negative cases with high confidence
- Patients are diagnosed as EP-negative, with high confidence
We take the same actions in the EP-positive case (admit patient and begin treatment) and the EP-negative case (discharge patient) as before, but now we have this middle ground. What should we do? Well, we can use resources to evaluate the patient more carefully. Perhaps there’s some kind of blood test \( T_2 \), which costs \$100, but improves our ability to distinguish between EP-positive and EP-negative populations. It’s more expensive than the \$1 test, but much less expensive than the \$5000 run of tests we used in false positive cases in our example.
How can we evaluate test \( T_2 \)?
Bivariate normal distributions
Let’s say that \( T_2 \) also has a numeric result \( y \) from 0 to 100, and it has a Gaussian distribution as well, so that tests \( T_1 \) and \( T_2 \) return a pair of values \( (x,y) \) with a bivariate normal distribution, in particular both \( x \) and \( y \) can be described by their mean values \( \mu_x, \mu_y \), standard deviations \( \sigma_x, \sigma_y \) and correlation coefficient \( \rho \), so that the covariance matrix \( \operatorname{cov}(x,y) = \begin{bmatrix}S_{xx} & S_{xy} \cr S_{xy} & S_{yy}\end{bmatrix} \) can be calculated with \( S_{xx} = \sigma_x{}^2, S_{yy} = \sigma_y{}^2, S_{xy} = \rho\sigma_x\sigma_y. \)
When a patient has EP (condition A), the second-order statistics of \( (x,y) \) can be described as
- \( \mu_x = 55, \mu_y = 57 \)
- \( \sigma_x=5.3, \sigma_y=4.1 \)
- \( \rho = 0.91 \)
When a patient does not have EP (condition B), the second-order statistics of \( (x,y) \) can be described as
- \( \mu_x = 40, \mu_y = 36 \)
- \( \sigma_x=5.9, \sigma_y=5.2 \)
- \( \rho = 0.84 \)
The covariance matrix may be unfamiliar to you, but it’s not very complicated. (Still, if you don’t like the math, just skip to the graphs below.) Each of the entries of the covariance matrix is merely the expected value of the product of the variables in question after removing the mean, so with a pair of zero-mean random variables \( (x’,y’) \) with \( x’ = x - \mu_x, y’=y-\mu_y \), the covariance matrix is just \( \operatorname{cov}(x’,y’) = \begin{bmatrix}E[x’^2] & E[x’y’] \cr E[x’y’] & E[y’^2] \end{bmatrix} \)
In order to help visualize this, let’s graph the two conditions.
First of all, we need to know how to generate pseudorandom values with these distributions. If we generate two independent Gaussian random variables \( (u,v) \) with zero mean and unit standard deviation, then the covariance matrix is just \( \begin{bmatrix}1 & 0 \cr 0 & 1\end{bmatrix} \). We can create new random variables \( (x,y) \) as a linear combination of \( u \) and \( v \):
$$\begin{aligned}x &= a_1u + b_1v \cr y &= a_2u + b_2v \end{aligned}$$
In this case, \( \operatorname{cov}(x,y)=\begin{bmatrix}E[x^2] & E[xy] \cr E[xy] & E[y^2] \end{bmatrix} = \begin{bmatrix}a_1{}^2 + b_1{}^2 & a_1a_2 + b_1b_2 \cr a_1a_2 + b_1b_2 & a_2{}^2 + b_2{}^2 \end{bmatrix}. \) As an example for computing this, \( E[x^2] = E[(a_1u+b_1v)^2] = a_1{}^2E[u^2] + 2a_1b_1E[uv] + b_1{}^2E[v^2] = a_1{}^2 + b_1{}^2 \) since \( E[u^2]=E[v^2] = 1 \) and \( E[uv]=0 \).
We can choose values \( a_1, a_2, b_1, b_2 \) so that we achieve the desired covariance matrix:
$$\begin{aligned} a_1 &= \sigma_x \cos \theta_x \cr b_1 &= \sigma_x \sin \theta_x \cr a_2 &= \sigma_y \cos \theta_y \cr b_2 &= \sigma_y \sin \theta_y \cr \end{aligned}$$
which yields \( \operatorname{cov}(x,y) = \begin{bmatrix}\sigma_x^2 & \sigma_x\sigma_y\cos(\theta_x -\theta_y) \cr \sigma_x\sigma_y\cos(\theta_x -\theta_y) & \sigma_y^2 \end{bmatrix}, \) and therefore we can choose any \( \theta_x, \theta_y \) such that \( \cos(\theta_x -\theta_y) = \rho. \) In particular, we can always choose \( \theta_x = 0 \) and \( \theta_y = \cos^{-1}\rho \), so that
$$\begin{aligned} a_1 &= \sigma_x \cr b_1 &= 0 \cr a_2 &= \sigma_y \rho \cr b_2 &= \sigma_y \sqrt{1-\rho^2} \cr \end{aligned}$$
and therefore
$$\begin{aligned} x &= \mu_x + \sigma_x u \cr y &= \mu_y + \rho\sigma_y u + \sqrt{1-\rho^2}\sigma_y v \end{aligned}$$
is a possible method of constructing \( (x,y) \) from independent unit Gaussian random variables \( (u,v). \) (For the mean values, we just added them in at the end.)
OK, so let’s use this to generate samples from the two conditions A and B, and graph them:
from scipy.stats import chi2 import matplotlib.colors colorconv = matplotlib.colors.ColorConverter() Coordinate2D = namedtuple('Coordinate2D','x y') class Gaussian2D(namedtuple('Gaussian','mu_x mu_y sigma_x sigma_y rho name id color')): @property def mu(self): """ mean """ return Coordinate2D(self.mu_x, self.mu_y) def cov(self): """ covariance matrix """ crossterm = self.rho*self.sigma_x*self.sigma_y return np.array([[self.sigma_x**2, crossterm], [crossterm, self.sigma_y**2]]) def sample(self, N, r=np.random): """ generate N random samples """ u = r.randn(N) v = r.randn(N) return self._transform(u,v) def _transform(self, u, v): """ transform from IID (u,v) to (x,y) with this distribution """ rhoc = np.sqrt(1-self.rho**2) x = self.mu_x + self.sigma_x*u y = self.mu_y + self.sigma_y*self.rho*u + self.sigma_y*rhoc*v return x,y def uv2xy(self, u, v): return self._transform(u,v) def xy2uv(self, x, y): rhoc = np.sqrt(1-self.rho**2) u = (x-self.mu_x)/self.sigma_x v = ((y-self.mu_y) - self.sigma_y*self.rho*u)/rhoc/self.sigma_y return u,v def contour(self, c, npoint=360): """ generate elliptical contours enclosing a fraction c of the population (c can be a vector) R^2 is a chi-squared distribution with 2 degrees of freedom: """ r = np.sqrt(chi2.ppf(c,2)) if np.size(c) > 1 and len(np.shape(c)) == 1: r = np.atleast_2d(r).T th = np.arange(npoint)*2*np.pi/npoint return self._transform(r*np.cos(th), r*np.sin(th)) def pdf_exponent(self, x, y): xdelta = x - self.mu_x ydelta = y - self.mu_y return -0.5/(1-self.rho**2)*( xdelta**2/self.sigma_x**2 - 2.0*self.rho*xdelta*ydelta/self.sigma_x/self.sigma_y + ydelta**2/self.sigma_y**2 ) @property def pdf_scale(self): return 1.0/2/np.pi/np.sqrt(1-self.rho**2)/self.sigma_x/self.sigma_y def pdf(self, x, y): """ probability density function """ q = self.pdf_exponent(x,y) return self.pdf_scale * np.exp(q) def logpdf(self, x, y): return np.log(self.pdf_scale) + self.pdf_exponent(x,y) @property def logpdf_coefficients(self): """ returns a vector (a,b,c,d,e,f) such that log(pdf(x,y)) = ax^2 + bxy + cy^2 + dx + ey + f """ f0 = np.log(self.pdf_scale) r = -0.5/(1-self.rho**2) a = r/self.sigma_x**2 b = r*(-2.0*self.rho/self.sigma_x/self.sigma_y) c = r/self.sigma_y**2 d = -2.0*a*self.mu_x -b*self.mu_y e = -2.0*c*self.mu_y -b*self.mu_x f = f0 + a*self.mu_x**2 + c*self.mu_y**2 + b*self.mu_x*self.mu_y return np.array([a,b,c,d,e,f]) def project(self, axis): """ Returns a 1-D distribution on the specified axis """ if isinstance(axis, basestring): if axis == 'x': mu = self.mu_x sigma = self.sigma_x elif axis == 'y': mu = self.mu_y sigma = self.sigma_y else: raise ValueError('axis must be x or y') else: # assume linear combination of x,y a,b = axis mu = a*self.mu_x + b*self.mu_y sigma = np.sqrt((a*self.sigma_x)**2 +(b*self.sigma_y)**2 +2*a*b*self.rho*self.sigma_x*self.sigma_y) return Gaussian1D(mu,sigma,self.name,self.id,self.color) def slice(self, x=None, y=None): """ Returns information (w, mu, sigma) on the probability distribution with x or y constrained: w: probability density across the entire slice mu: mean value of the pdf within the slice sigma: standard deviation of the pdf within the slice """ if x is None and y is None: raise ValueError("At least one of x or y must be a value") rhoc = np.sqrt(1-self.rho**2) if y is None: w = scipy.stats.norm.pdf(x, self.mu_x, self.sigma_x) mu = self.mu_y + self.rho*self.sigma_y/self.sigma_x*(x-self.mu_x) sigma = self.sigma_y*rhoc else: # x is None w = scipy.stats.norm.pdf(y, self.mu_y, self.sigma_y) mu = self.mu_x + self.rho*self.sigma_x/self.sigma_y*(y-self.mu_y) sigma = self.sigma_x*rhoc return w, mu, sigma def slicefunc(self, which): rhoc = np.sqrt(1-self.rho**2) if which == 'x': sigma = self.sigma_y*rhoc a = self.rho*self.sigma_y/self.sigma_x def f(x): w = scipy.stats.norm.pdf(x, self.mu_x, self.sigma_x) mu = self.mu_y + a*(x-self.mu_x) return w,mu,sigma elif which == 'y': sigma = self.sigma_x*rhoc a = self.rho*self.sigma_x/self.sigma_y def f(y): w = scipy.stats.norm.pdf(y, self.mu_y, self.sigma_y) mu = self.mu_x + a*(y-self.mu_y) return w,mu,sigma else: raise ValueError("'which' must be x or y") return f DETERMINISTIC_SEED = 123 np.random.seed(DETERMINISTIC_SEED) N = 100000 distA = Gaussian2D(mu_x=55,mu_y=57,sigma_x=5.3,sigma_y=4.1,rho=0.91,name='A (EP-positive)',id='A',color='red') distB = Gaussian2D(mu_x=40,mu_y=36,sigma_x=5.9,sigma_y=5.2,rho=0.84,name='B (EP-negative)',id='B',color='#8888ff') xA,yA = distA.sample(N) xB,yB = distB.sample(N) fig=plt.figure() ax=fig.add_subplot(1,1,1) def scatter_samples(ax,xyd_list,contour_list=(),**kwargs): Kmute = 1 if not contour_list else 0.5 for x,y,dist in xyd_list: mutedcolor = colorconv.to_rgb(dist.color) mutedcolor = [c*Kmute+(1-Kmute) for c in mutedcolor] if not contour_list: kwargs['label']=dist.name ax.plot(x,y,'.',color=mutedcolor,alpha=0.8,markersize=0.5,**kwargs) for x,y,dist in xyd_list: # Now draw contours for certain probabilities th = np.arange(1200)/1200.0*2*np.pi u = np.cos(th) v = np.sin(th) first = True for p in contour_list: cx,cy = dist.contour(p) kwargs = {} if first: kwargs['label']=dist.name first = False ax.plot(cx,cy,color=dist.color,linewidth=0.5,**kwargs) ax.set_xlabel('x') ax.set_ylabel('y') ax.legend(loc='lower right',markerscale=10) ax.grid(True) title = 'Scatter sample plot' if contour_list: title += (', %s%% CDF ellipsoid contours shown' % (', '.join('%.0f' % (p*100) for p in contour_list))) ax.set_title(title, fontsize=10) scatter_samples(ax,[(xA,yA,distA), (xB,yB,distB)], [0.25,0.50,0.75,0.90,0.95,0.99]) for x,y,desc in [(xA, yA,'A'),(xB,yB,'B')]: print "Covariance matrix for case %s:" % desc C = np.cov(x,y) print C sx = np.sqrt(C[0,0]) sy = np.sqrt(C[1,1]) rho = C[0,1]/sx/sy print "sample sigma_x = %.3f" % sx print "sample sigma_y = %.3f" % sy print "sample rho = %.3f" % rho
Covariance matrix for case A: [[ 28.06613839 19.74199382] [ 19.74199382 16.76597717]] sample sigma_x = 5.298 sample sigma_y = 4.095 sample rho = 0.910 Covariance matrix for case B: [[ 34.6168817 25.69386711] [ 25.69386711 27.05651845]] sample sigma_x = 5.884 sample sigma_y = 5.202 sample rho = 0.840
It may seem strange, but having the results of two tests (\( x \) from test \( T_1 \) and \( y \) from test \( T_2 \)) gives more useful information than the result of each test considered on its own. We’ll come back to this idea a little bit later.
The more immediate question is: given the pair of results \( (x,y) \), how would we decide whether the patient has \( EP \) or not? With just test \( T_1 \) we could merely declare an EP-positive diagnosis if \( x > x_0 \) for some threshold \( x_0 \). With two variables, some kind of inequality is involved, but how do we decide?
Bayes’ Rule
We are greatly indebted to the various European heads of state and religion during much of the 18th century (the Age of Enlightenment) for merely leaving people alone. (OK, this wasn’t universally true, but many of the monarchies turned a blind eye towards intellectualism.) This lack of interference and oppression resulted in numerous mathematical and scientific discoveries, one of which was Bayes’ Rule, named after Thomas Bayes, a British clergyman and mathematician. Bayes’ Rule was published posthumously in An Essay towards solving a Problem in the Doctrine of Chances, and later inflicted on throngs of undergraduate students of probability and statistics.
The basic idea involves conditional probabilities and reminds me of the logical converse. As a hypothetical example, suppose we know that 95% of Dairy Queen customers are from the United States and that 45% of those US residents who visit Dairy Queen like peppermint ice cream, whereas 72% of non-US residents like peppermint ice cream. We are in line to get some ice cream, and we notice that the person in front of us orders peppermint ice cream. Can we make any prediction of the probability that this person is from the US?
Bayes’ Rule relates these two conditions. Let \( A \) represent the condition that a Dairy Queen customer is a US resident, and \( B \) represent that they like peppermint ice cream. Then \( P(A\ |\ B) = \frac{P(B\ |\ A)P(A)}{P(B)} \), which is really just an algebraic rearrangement of the expression of the joint probability that \( A \) and \( B \) are both true: \( P(AB) = P(A\ |\ B)P(B) = P(B\ |\ A)P(A) \). Applied to our Dairy Queen example, we have \( P(A) = 0.95 \) (95% of Dairy Queen customers are from the US) and \( P(B\ |\ A) = 0.45 \) (45% of customers like peppermint ice cream, given that they are from the US). But what is \( P(B) \), the probability that a Dairy Queen customer likes peppermint ice cream? Well, it’s the sum of the all the constituent joint probabilities where the customer likes peppermint ice cream. For example, \( P(AB) = P(B\ |\ A)P(A) = 0.45 \times 0.95 = 0.4275 \) is the joint probability that a Dairy Queen customer is from the US and likes peppermint ice cream, and \( P(\bar{A}B) = P(B\ |\ \bar{A})P(\bar{A}) = 0.72 \times 0.05 = 0.036 \) is the joint probability that a Dairy Queen customer is not from the US and likes peppermint ice cream. Then \( P(B) = P(AB) + P(\bar{A}B) = 0.4275 + 0.036 = 0.4635 \). (46.35% of all Dairy Queen customers like peppermint ice cream.) The final application of Bayes’ Rule tells us
$$P(A\ |\ B) = \frac{P(B\ |\ A)P(A)}{P(B)} = \frac{0.45 \times 0.95}{0.4635} \approx 0.9223$$
and therefore if we see someone order peppermint ice cream at Dairy Queen, there is a whopping 92.23% chance they are not from the US.
Let’s go back to our embolary pulmonism scenario where a person takes both tests \( T_1 \) and \( T_2 \), with results \( R = (x=45, y=52) \). Can we estimate the probability that this person has EP?
N = 500000 np.random.seed(DETERMINISTIC_SEED) xA,yA = distA.sample(N) xB,yB = distB.sample(N) fig=plt.figure() ax=fig.add_subplot(1,1,1) scatter_samples(ax,[(xA,yA,distA), (xB,yB,distB)]) ax.plot(45,52,'.k');
We certainly aren’t going to be able to find the answer exactly just from looking at this chart, but it looks like an almost certain case of A being true — that is, \( R: x=45, y=52 \) implies that the patient probably has EP. Let’s figure it out as
$$P(A\ |\ R) = \frac{P(R\ |\ A)P(A)}{P(R)}.$$
Remember we said earlier that the base rate, which is the probability of any given person presenting symptoms having EP before any testing, is \( P(A) = 40\times 10^{-6} \). (This is known as the a priori probability, whenever this Bayesian stuff gets involved.) The other two probabilities \( P(R\ |\ A) \) and \( P(R) \) are technically infinitesimal, because they are part of continuous probability distributions, but we can handwave and say that \( R \) is really the condition that the results are \( 45 \le x \le 45 + dx \) and \( 52 \le y \le 52+dy \) for some infinitesimal interval widths \( dx, dy \), in which case \( P(R\ |\ A) = p_A(R)\,dx\,dy \) and \( P(R) = P(R\ |\ A)P(A) + P(R\ |\ B)P(B) = p_A(R)P(A)\,dx\,dy + p_B(R)P(B)\,dx\,dy \) where \( p_A \) and \( p_B \) are the probability density functions. Substituting this all in we get
$$P(A\ |\ R) = \frac{p_A(R)P(A)}{p_A(R)P(A)+p_B(R)P(B)}$$
The form of the bivariate normal distribution is not too complicated, just a bit unwieldy:
$$p(x,y) = \frac{1}{2\pi\sqrt{1-\rho^2}\sigma_x\sigma_y}e^{q(x,y)}$$
with
$$q(x,y) = -\frac{1}{2(1-\rho^2)}\left(\frac{(x-\mu_x)^2}{\sigma_x{}^2}-2\rho\frac{(x-\mu_x)(y-\mu_y)}{\sigma_x\sigma_y}+\frac{(y-\mu_y)^2}{\sigma_y{}^2}\right)$$
and the rest is just number crunching:
x1=45 y1=52 PA_total = 40e-6 pA = distA.pdf(x1, y1) pB = distB.pdf(x1, y1) print "pA(%.1f,%.1f) = %.5g" % (x1,y1,pA) print "pB(%.1f,%.1f) = %.5g" % (x1,y1,pB) print ("Bayes' rule result: p(A | x=%.1f, y=%.1f) = %.5g" % (x1,y1,pA*PA_total/(pA*PA_total+pB*(1-PA_total))))
pA(45.0,52.0) = 0.0014503 pB(45.0,52.0) = 4.9983e-07 Bayes' rule result: p(A | x=45.0, y=52.0) = 0.104
Wow, that’s counterintuitive. This result value \( R \) lies much closer to the probability “cloud” of A = EP-positive than B = EP-negative, but Bayes’ Rule tells us there’s only about a 10.4% probability that a patient with test results \( (x=45, y=52) \) has EP. And it’s because of the very low incidence of EP.
There is something we can do to make reading the graph useful, and that’s to plot a parameter I’m going to call \( \lambda(x,y) \), which is the logarithm of the ratio of probability densities:
$$\lambda(x,y) = \ln \frac{p_A(x,y)}{p_B(x,y)} = \ln p_A(x,y) - \ln p_B(x,y)$$
Actually, we’ll plot \( \lambda_{10}(x,y) = \lambda(x,y) / \ln 10 = \log_{10} \frac{p_A(x,y)}{p_B(x,y)} \).
This parameter is useful because Bayes’ rule calculates
$$\begin{aligned} P(A\ |\ x,y) &= \frac{p_A(x,y) P_A} { p_A(x,y) P_A + p_B(x,y) P_B} \cr &= \frac{p_A(x,y)/p_B(x,y) P_A} {p_A(x,y)/p_B(x,y) P_A + P_B} \cr &= \frac{e^{\lambda(x,y)} P_A} {e^{\lambda(x,y)} P_A + P_B} \cr &= \frac{1}{1 + e^{-\lambda(x,y)}P_B / P_A} \end{aligned}$$
and for any desired \( P(A\ |\ x,y) \) we can figure out the equivalent value of \( \lambda(x,y) = - \ln \left(\frac{P_A}{P_B}((1/P(A\ |\ x,y) - 1)\right). \)
This means that for a fixed value of \( \lambda \), then \( P(A\ |\ \lambda) = \frac{1}{1 + e^{-\lambda}P_B / P_A}. \)
For bivariate Gaussian distributions, the \( \lambda \) parameter is also useful because it is a quadratic function of \( x \) and \( y \), so curves of constant \( \lambda \) are conic sections (lines, ellipses, hyperbolas, or parabolas).
def jcontourp(ax,x,y,z,levels,majorfunc,color=None,fmt=None, **kwargs): linewidths = [1 if majorfunc(l) else 0.4 for l in levels] cs = ax.contour(x,y,z,levels, linewidths=linewidths, linestyles='-', colors=color,**kwargs) labeled = [l for l in cs.levels if majorfunc(l)] ax.clabel(cs, labeled, inline=True, fmt='%s', fontsize=10) xv = np.arange(10,80.01,0.1) yv = np.arange(10,80.01,0.1) x,y = np.meshgrid(xv,yv) def lambda10(distA,distB,x,y): return (distA.logpdf(x, y)-distB.logpdf(x, y))/np.log(10) fig = plt.figure() ax = fig.add_subplot(1,1,1) scatter_samples(ax,[(xA,yA,distA), (xB,yB,distB)], zorder=-1) ax.plot(x1,y1,'.k') print "lambda10(x=%.1f,y=%.1f) = %.2f" % (x1,y1,lambda10(distA,distB,x1,y1)) levels = np.union1d(np.arange(-10,10), np.arange(-200,100,10)) def levelmajorfunc(level): if -10 <= level <= 10: return int(level) % 5 == 0 else: return int(level) % 25 == 0 jcontourp(ax,x,y,lambda10(distA,distB,x,y), levels, levelmajorfunc, color='black') ax.set_xlim(xv.min(), xv.max()+0.001) ax.set_ylim(yv.min(), yv.max()+0.001) ax.set_title('Scatter sample plot with contours = $\lambda_{10}$ values');
lambda10(x=45.0,y=52.0) = 3.46
We’ll pick one particular \( \lambda_{10} \) value as a threshold \( L_{10} = L/ \ln 10 \), and if \( \lambda_{10} > L_{10} \) then we’ll declare condition \( a \) (the patient is diagnosed as EP-positive), otherwise we’ll declare condition \( b \) (the patient is diagnosed as EP-negative). The best choice of \( L_{10} \) is the one that maximizes expected value.
Remember how we did this in the case with only the one test \( T_1 \) with result \( x \): we chose threshold \( x_0 \) based on the point where \( {\partial E \over \partial x_0} = 0 \); in other words, a change in diagnosis did not change the expected value at this point. We can do the same thing here:
$$0 = {\partial E[v] \over \partial {L}} = P(A\ |\ \lambda=L)(v_{Ab}-v_{Aa}) + P(B\ |\ \lambda=L)(v_{Bb}-v_{Ba})$$
With our earlier definitions
$$\rho_v = -\frac{v_{Bb}-v_{Ba}}{v_{Ab}-v_{Aa}}, \qquad \rho_p = P_B / P_A$$
then the equation for \( L \) becomes \( 0 = P(A\ |\ \lambda=L) - \rho_v P(B\ |\ \lambda=L) = P(A\ |\ \lambda=L) - \rho_v (1-P(A\ |\ \lambda=L)) \), which simplifies to
$$P(A\ |\ \lambda=L) = \frac{\rho_v}{\rho_v+1} = \frac{1}{1+1/\rho_v} .$$
But we already defined \( \lambda \) as the group of points \( (x,y) \) that have equal probability \( P(A\ |\ \lambda) = \frac{1}{1 + e^{-\lambda}P_B / P_A} = \frac{1}{1 + \rho_p e^{-\lambda}} \) so
$$\frac{1}{1+1/\rho_v} = \frac{1}{1 + \rho_p e^{-L}}$$
which occurs when \( L = \ln \rho_v \rho_p. \)
In our EP example,
$$\begin{aligned} \rho_p &= p_B/p_A = 0.99996 / 0.00004 = 24999 \cr \rho_v &= -\frac{v_{Bb}-v_{Ba}}{v_{Ab}-v_{Aa}} = -5000 / -9.9\times 10^6 \approx 0.00050505 \cr \rho_v\rho_p &\approx 12.6258 \cr L &= \ln \rho_v\rho_p \approx 2.5357 \cr L_{10} &= \log_{10}\rho_v\rho_p \approx 1.1013 \end{aligned}$$
and we can complete the analysis empirically by looking at the fraction of pseudorandomly-generated sample points where \( \lambda_{10} < L_{10} \); this is an example of Monte Carlo analysis.
class Quadratic1D(namedtuple('Quadratic1D','a b c')): """ Q(x) = a*x*x + b*x + c """ __slots__ = () @property def x0(self): return -self.b/(2.0*self.a) @property def q0(self): return self.c - self.a*self.x0**2 def __call__(self,x): return self.a*x*x + self.b*x + self.c def solve(self, q): D = self.b*self.b - 4*self.a*(self.c-q) sqrtD = np.sqrt(D) return np.array([-self.b-sqrtD, -self.b+sqrtD])/(2*self.a) class QuadraticLissajous(namedtuple('QuadraticLissajous','x0 y0 Rx Ry phi')): """ A parametric curve as a function of theta: x = x0 + Rx * cos(theta) y = y0 + Ry * sin(theta+phi) """ __slots__ = () def __call__(self, theta): return (self.x0 + self.Rx * np.cos(theta), self.y0 + self.Ry * np.sin(theta+self.phi)) class Quadratic2D(namedtuple('Quadratic2D','a b c d e f')): """ Bivariate quadratic function Q(x,y) = a*x*x + b*x*y + c*y*y + d*x + e*y + f = a*(x-x0)*(x-x0) + b*(x-x0)*(y-y0) + c*(y-y0)*(y-y0) + q0 = s*(u*u + v*v) + q0 where s = +/-1 (Warning: this implementation assumes convexity, that is, b*b < 4*a*c, so hyperboloids/paraboloids are not handled.) """ __slots__ = () @property def discriminant(self): return self.b**2 - 4*self.a*self.c @property def x0(self): return (2*self.c*self.d - self.b*self.e)/self.discriminant @property def y0(self): return (2*self.a*self.e - self.b*self.d)/self.discriminant @property def q0(self): x0 = self.x0 y0 = self.y0 return self.f - self.a*x0*x0 - self.b*x0*y0 - self.c*y0*y0 def _Kcomponents(self): s = 1 if self.a > 0 else -1 r = s*self.b/2.0/np.sqrt(self.a*self.c) rc = np.sqrt(1-r*r) Kux = rc*np.sqrt(self.a*s) Kvx = r*np.sqrt(self.a*s) Kvy = np.sqrt(self.c*s) return Kux, Kvx, Kvy @property def Kxy2uv(self): Kux, Kvx, Kvy = self._Kcomponents() return np.array([[Kux, 0],[Kvx, Kvy]]) @property def Kuv2xy(self): Kux, Kvx, Kvy = self._Kcomponents() return np.array([[1.0/Kux, 0],[-1.0*Kvx/Kux/Kvy, 1.0/Kvy]]) @property def transform_xy2uv(self): Kxy2uv = self.Kxy2uv x0 = self.x0 y0 = self.y0 def transform(x,y): return np.dot(Kxy2uv, [x-x0,y-y0]) return transform @property def transform_uv2xy(self): Kuv2xy = self.Kuv2xy x0 = self.x0 y0 = self.y0 def transform(u,v): return np.dot(Kuv2xy, [u,v]) + [[x0],[y0]] return transform def uv_radius(self, q): """ Returns R such that solutions (u,v) of Q(x,y) = q lie within the range [-R, R], or None if there are no solutions. """ s = 1 if self.a > 0 else -1 D = (q-self.q0)*s return np.sqrt(D) if D >= 0 else None def _xy_radius_helper(self, q, z): D = (self.q0 - q) * 4 * z / self.discriminant if D < 0: return None else: return np.sqrt(D) def x.c) def y.a) def lissajous(self, q): """ Returns a QuadraticLissajous with x0, y0, Rx, Ry, phi such that the solutions (x,y) of Q(x,y) = q can be written: x = x0 + Rx * cos(theta) y = y0 + Ry * sin(theta+phi) Rx and Ry and phi may each return None if no such solution exists. """ D = self.discriminant x0 = (2*self.c*self.d - self.b*self.e)/D y0 = (2*self.a*self.e - self.b*self.d)/D q0 = self.f - self.a*x0*x0 - self.b*x0*y0 - self.c*y0*y0 Dx = 4 * (q0-q) * self.c / D Rx = None if Dx < 0 else np.sqrt(Dx) Dy = 4 * (q0-q) * self.a / D Ry = None if Dy < 0 else np.sqrt(Dy) phi = None if D > 0 else np.arcsin(self.b / (2*np.sqrt(self.a*self.c))) return QuadraticLissajous(x0,y0,Rx,Ry,phi) def contour(self, q, npoints=360): """ Returns a pair of arrays x,y such that Q(x,y) = q """ s = 1 if self.a > 0 else -1 R = np.sqrt((q-self.q0)*s) th = np.arange(npoints)*2*np.pi/npoints u = R*np.cos(th) v = R*np.sin(th) return self.transform_uv2xy(u,v) def constrain(self, x=None, y=None): if x is None and y is None: return self if x is None: # return a function in x return Quadratic1D(self.a, self.d + y*self.b, self.f + y*self.e + y*y*self.c) if y is None: # return a function in y return Quadratic1D(self.c, self.e + x*self.b, self.f + x*self.d + x*x*self.a) return self(x,y) def __call__(self, x, y): return (self.a*x*x +self.b*x*y +self.c*y*y +self.d*x +self.e*y +self.f) def decide_limits(*args, **kwargs): s = kwargs.get('s', 6) xmin = None xmax = None for xbatch in args: xminb = min(xbatch) xmaxb = max(xbatch) mu = np.mean(xbatch) std = np.std(xbatch) xminb = min(mu-s*std, xminb) xmaxb = max(mu+s*std, xmaxb) if xmin is None: xmin = xminb xmax = xmaxb else: xmin = min(xmin,xminb) xmax = max(xmax,xmaxb) # Quantization q = kwargs.get('q') if q is not None: xmin = np.floor(xmin/q)*q xmax = np.ceil(xmax/q)*q return xmin, xmax def separation_plot(xydistA, xydistB, Q, L, ax=None, xlim=None, ylim=None): L10 = L/np.log(10) if ax is None: fig = plt.figure() ax = fig.add_subplot(1,1,1) xA,yA,distA = xydistA xB,yB,distB = xydistB scatter_samples(ax,[(xA,yA,distA), (xB,yB,distB)], zorder=-1) xc,yc = Q.contour(L) ax.plot(xc,yc, color='green', linewidth=1.5, dashes=[5,2], label = '$\\lambda_{10} = %.4f$' % L10) ax.legend(loc='lower right',markerscale=10, labelspacing=0,fontsize=12) if xlim is None: xlim = decide_limits(xA,xB,s=6,q=10) if ylim is None: ylim = decide_limits(yA,yB,s=6,q=10) xv = np.arange(xlim[0],xlim[1],0.1) yv = np.arange(ylim[0],ylim[1],0.1) x,y = np.meshgrid(xv,yv) jcontourp(ax,x,y,lambda10(distA,distB,x,y), levels, levelmajorfunc, color='black') ax.set_xlim(xlim) ax.set_ylim(ylim) ax.set_title('Scatter sample plot with contours = $\lambda_{10}$ values') def separation_report(xydistA, xydistB, Q, L): L10 = L/np.log(10) for x,y,dist in [xydistA, xydistB]: print "Separation of samples in %s by L10=%.4f" % (dist.id,L10) lam = Q(x,y) lam10 = lam/np.log(10) print " Range of lambda10: %.4f to %.4f" % (np.min(lam10), np.max(lam10)) n = np.size(lam) p = np.count_nonzero(lam < L) * 1.0 / n print " lambda10 < L10: %.5f" % p print " lambda10 >= L10: %.5f" % (1-p) L = np.log(5000 / 9.9e6 * 24999) C = distA.logpdf_coefficients - distB.logpdf_coefficients Q = Quadratic2D(*C) separation_plot((xA,yA,distA),(xB,yB,distB), Q, L) separation_report((xA,yA,distA),(xB,yB,distB), Q, L)
Or we can determine the results by numerical integration of probability density.
The math below isn’t difficult, just tedious; for each of the two Gaussian distributions for A and B, I selected a series of 5000 intervals (10001 x-axis points) from \( \mu_x-8\sigma_x \) to \( \mu_x + 8\sigma_x \), and used Simpson’s Rule to integrate the probability density \( f_x(x_i) \) at each point \( x_i \), given that
$$x \approx x_i \quad\Rightarrow\quad f_x(x) \approx f_x(x_i) = p_x(x_i) \left(F_{x_i}(y_{i2}) - F_{x_i}(y_{i1})\right)$$
where
- \( p_x(x_i) \) is the probability density that \( x=x_i \) with \( y \) unspecified
- \( F_{x_i}(y_0) \) is the 1-D cumulative distribution function, given \( x=x_i \), that \( y<y_0 \)
- \( y_{i1} \) and \( y_{i2} \) are either
- the two solutions of \( \lambda(x_i,y) = L \)
- or zero, if there are no solutions or one solution
and the sample points \( x_i \) are placed more closely together nearer to the extremes of the contour \( \lambda(x,y)=L \) to capture the suddenness of the change.
Anyway, here’s the result, which is nicely consistent with the Monte Carlo analysis:
def assert_ordered(*args, **kwargs): if len(args) < 2: return rthresh = kwargs.get('rthresh',1e-10) label = kwargs.get('label',None) label = '' if label is None else label+': ' xmin = args[0] xmax = args[-1] if len(args) == 2: # not very interesting case xthresh = rthresh*(xmax+xmin)/2.0 else: xthresh = rthresh*(xmax-xmin) xprev = xmin for x in args[1:]: assert x - xprev >= -xthresh, "%s%s > %s + %g" % (label,xprev,x,xthresh) xprev = x def arccos_sat(x): if x <= -1: return np.pi if x >= 1: return 0 return np.arccos(x) def simpsons_rule_points(xlist, bisect=True): """ Generator for Simpson's rule xlist: arbitrary points in increasing order bisect: whether or not to add bisection points returns a generator of weights w (if bisect=False) or tuples (w,x) (if bisect = True) such that the integral of f(x) dx over the list of points xlist is approximately equal to: sum(w*f(x) for w,x in simpsons_rule_points(xlist)) The values of x returned are x[i], (x[i]+x[i+1])/2, x[i+1] with relative weights dx/6, 4*dx/6, dx/6 for each interval [x[i], x[i+1]] """ xiter = iter(xlist) xprev = xiter.next() w2 = 0 x2 = None if bisect: for x2 in xiter: x0 = xprev dx = x2-x0 x1 = x0 + dx/2.0 xprev = x2 w6 = dx/6.0 w0 = w2 + w6 yield (w0, x0) w1 = 4*w6 yield (w1, x1) w2 = w6 if x2 is not None: yield (w2, x2) else: for x1 in xiter: x0 = xprev try: x2 = xiter.next() except StopIteration: raise ValueError("Must have an odd number of points") dx = x2-x0 xprev = x2 w6 = dx/6.0 w0 = w2 + w6 yield w0 w1 = 4*w6 yield w1 w2 = w6 if x2 is not None: yield w2 def estimate_separation_numerical(dist, Q, L, xmin, xmax, Nintervals=5000, return_pair=False, sampling_points=None): """ Numerical estimate of each row of confusion matrix using N integration intervals (each bisected to use Simpson's Rule) covering the interval [xmin,xmax]. dist: Gaussian2D distribution Q: Quadratic2D function Q(x,y) L: lambda threshold: diagnose as A when Q(x,y) > L, otherwise as B xmin: start x-coordinate xmax: end x-coordinate Nintervals: number of integration intervals Returns p, where p is the probability of Q(x,y) > L for that distribution. If return_pair is true, returns [p,q]; if [x1,x2] covers most of the distribution, then p+q should be close to 1. """ Qliss = Q.lissajous(L) xqr = Qliss.Rx xqmu = Qliss.x0 # Determine range of solutions: # no solutions if xqr is None, # otherwise Q(x,y) > L if x in Rx = [xqmu-xqr, xqmu+xqr] # # Cover the trivial cases first: if xqr is None: return (0,1) if return_pair else 0 if (xmax < xqmu - xqr) or (xmin > xqmu + xqr): # All solutions are more than Nsigma from the mean # isigma_left = (xqmu - xqr - dist.mu_x)/dist.sigma_x # isigma_right = (xqmu + xqr - dist.mu_x)/dist.sigma_x # print "sigma", isigma_left, isigma_right return (0,1) if return_pair else 0 # We want to cover the range xmin, xmax. # Increase coverage near the ends of the lambda threshold, # xqmu +/- xqr where the behavior changes more rapidly, # by using points at -cos(theta) within the solution space for Q(x,y)>L th1 = arccos_sat(-(xmin-xqmu)/xqr) th2 = arccos_sat(-(xmax-xqmu)/xqr) # print np.array([th1,th2])*180/np.pi, xmin, xqmu-np.cos(th1)*xqr, xqmu-np.cos(th2)*xqr, xmax, (xqmu-xqr, xqmu+xqr) assert_ordered(xmin, xqmu-np.cos(th1)*xqr, xqmu-np.cos(th2)*xqr, xmax) x_arc_coverage = xqr*(np.cos(th1)-np.cos(th2)) # length along x x_arc_length = xqr*(th2-th1) # length along arc x_total_length = (xmax-xmin) - x_arc_coverage + x_arc_length x1 = xqmu-xqr*np.cos(th1) x2 = x1 + x_arc_length n = (Nintervals*2) + 1 xlin = np.linspace(0,x_total_length, n) + xmin x = xlin[:] y = np.zeros((2,n)) # Points along arc: tol = 1e-10 i12 = (xlin >= x1 - tol) & (xlin <= x2 + tol) angles = th1 + (xlin[i12]-x1)/x_arc_length*(th2-th1) x[i12], y[0, i12] = Qliss(np.pi + angles) _, y[1, i12] = Qliss(np.pi - angles) y.sort(axis=0) x[xlin >= x2] += (x_arc_coverage-x_arc_length) fx = dist.slicefunc('x') p = 0 q = 0 for i,wdx in enumerate(simpsons_rule_points(x, bisect=False)): w, mu, sigma = fx(x[i]) y1 = y[0,i] y2 = y[1,i] cdf1 = scipy.stats.norm.cdf(y1, mu, sigma) cdf2 = scipy.stats.norm.cdf(y2, mu, sigma) deltacdf = cdf2-cdf1 p += wdx*w*deltacdf q += wdx*w*(1-deltacdf) return (p,q) if return_pair else p def compute_confusion_matrix(distA, distB, Q, L, Nintervals=5000, verbose=False): C = np.zeros((2,2)) for i,dist in enumerate([distA,distB]): Nsigma = 8 xmin = dist.mu_x - Nsigma*dist.sigma_x xmax = dist.mu_x + Nsigma*dist.sigma_x p,q = estimate_separation_numerical(dist, Q, L, xmin, xmax, Nintervals=Nintervals, return_pair=True) C[i,:] = [p,q] print "%s: p=%g, q=%g, p+q-1=%+g" % (dist.name, p,q,p+q-1) return C confusion_matrix = compute_confusion_matrix(distA, distB, Q, L, verbose=True) separation_report((xA,yA,distA),(xB,yB,distB), Q, L) value_matrix = np.array([[0,-5000],[-1e7, -1e5]]) - 100 # Same value matrix as the original one # (with vAb = $10 million) before but we add $100 cost for test T2 C = confusion_matrix[::-1,::-1] # flip the confusion matrix left-right and top-bottom for B first def gather_info(C,V,PA,**kwargs): info = dict(kwargs) C = np.array(C) V = np.array(V) info['C'] = C info['V'] = V info['J'] = C*[[1-PA],[PA]] return info display(show_binary_matrix(C, 'L_{10}=%.4f' % (L/np.log(10)), [distB, distA], 'ba', 40e-6,value_matrix, special_format={'v':'$%.2f'})) info27 = gather_info(C,value_matrix,40e-6,L=L)
A (EP-positive): p=0.982467, q=0.0175334, p+q-1=-8.90805e-09 B (EP-negative): p=0.00163396, q=0.998366, p+q-1=+1.97626e-07
Report for threshold \(L_{10}=1.1013 \rightarrow E[v]=\)$-119.11
There! Now we can check that we have a local maximum, by trying slightly lower and higher thresholds:
value_matrix = np.array([[0,-5000],[-1e7, -1e5]]) - 100 delta_L = 0.1*np.log(10) for Li in [L-delta_L, L+delta_L]: confusion_matrix = compute_confusion_matrix(distA, distB, Q, Li, verbose=True) separation_report((xA,yA,distA),(xB,yB,distB), Q, Li) # Same value matrix before but we add $100 cost for test T2 C = confusion_matrix[::-1,::-1] # flip the confusion matrix left-right and top-bottom for B first display(show_binary_matrix(C, 'L_{10}=%.4f' % (Li/np.log(10)), [distB, distA], 'ba', 40e-6,value_matrix, special_format={'v':'$%.2f'}))
A (EP-positive): p=0.984803, q=0.0151974, p+q-1=-8.85844e-09 B (EP-negative): p=0.0018415, q=0.998159, p+q-1=+2.55256e-07 Separation of samples in A by L10=1.0013 Range of lambda10: -3.4263 to 8.1128 lambda10 < L10: 0.01550 lambda10 >= L10: 0.98450 Separation of samples in B by L10=1.0013 Range of lambda10: -36.8975 to 4.0556 lambda10 < L10: 0.99819 lambda10 >= L10: 0.00181
Report for threshold \(L_{10}=1.0013 \rightarrow E[v]=\)$-119.23
A (EP-positive): p=0.979808, q=0.0201915, p+q-1=-9.01419e-09 B (EP-negative): p=0.00144637, q=0.998554, p+q-1=+2.1961e-08 Separation of samples in A by L10=1.2013 Range of lambda10: -3.4263 to 8.1128 lambda10 < L10: 0.02011 lambda10 >= L10: 0.97989 Separation of samples in B by L10=1.2013 Range of lambda10: -36.8975 to 4.0556 lambda10 < L10: 0.99850 lambda10 >= L10: 0.00150
Report for threshold \(L_{10}=1.2013 \rightarrow E[v]=\)$-119.23
Great! The sensitivity of threshold seems pretty flat; \( E[v] \) differs by about 12 cents if we change \( L_{10} = 1.1013 \) to \( L_{10} = 1.0013 \) or \( L_{10} = 1.2013 \). This gives us a little wiggle room in the end to shift costs between the false-negative and false-positive cases, without changing the overall expected cost very much.
Most notably, though, we’ve reduced the total cost by about \$93 by using the pair of tests \( T_1, T_2 \) compared to the test with just \( T_1 \). This is because we shift cost from the Ab (false negative for EP-positive) and Ba (false positive for EP-negative) cases to the Bb (correct for EP-negative) case — everyone has to pay \$100 more, but the chances of false positives and false negatives have been greatly reduced.
Don’t remember the statistics from the one-test case? Here they are again:
x0 = 53.1626 C = analyze_binary(dneg, dpos, x0) value_matrix_T1 = [[0,-5000],[-1e7, -1e5]] display(show_binary_matrix(C, x0, [dneg, dpos], 'ba', 40e-6,value_matrix_T1, special_format={'v':'$%.2f'})) info17 = gather_info(C,value_matrix_T1,40e-6,x0=x0)
Report for threshold \(x_0 = 53.1626 \rightarrow E[v]=\)$-212.52
And the equivalent cases for \( v_{Ab}= \)\$100 million:
Pa_total = 40e-6 x0 = 48.1458 C1 = analyze_binary(dneg, dpos, x0) value_matrix_T1 = np.array([[0,-5000],[-1e8, -1e5]]) display(show_binary_matrix(C1, x0, [dneg, dpos], 'ba', Pa_total, value_matrix_T1, special_format={'v':'$%.2f'})) info18 = gather_info(C1,value_matrix_T1,Pa_total,x0=x0) def compute_optimal_L(value_matrix,Pa_total): v = value_matrix return np.log(-(v[0,0]-v[0,1])/(v[1,0]-v[1,1])*(1-Pa_total)/Pa_total) value_matrix_T2 = value_matrix_T1 - 100 L = compute_optimal_L(value_matrix_T2, Pa_total) confusion_matrix = compute_confusion_matrix(distA, distB, Q, L, verbose=True) separation_report((xA,yA,distA),(xB,yB,distB), Q, L) # Same value matrix before but we add $100 cost for test T2 C2 = confusion_matrix[::-1,::-1] # flip the confusion matrix left-right and top-bottom for B first display(show_binary_matrix(C2, 'L_{10}=%.4f' % (L/np.log(10)), [distB, distA], 'ba', Pa_total,value_matrix_T2, special_format={'v':'$%.2f'})) separation_plot((xA,yA,distA),(xB,yB,distB), Q, L) info28 = gather_info(C2,value_matrix_T2,Pa_total,L=L)
Report for threshold \(x_0 = 48.1458 \rightarrow E[v]=\)$-813.91
A (EP-positive): p=0.996138, q=0.00386181, p+q-1=-8.37725e-09 B (EP-negative): p=0.00491854, q=0.995081, p+q-1=+4.32711e-08 Separation of samples in A by L10=0.0973 Range of lambda10: -3.4263 to 8.1128 lambda10 < L10: 0.00389 lambda10 >= L10: 0.99611 Separation of samples in B by L10=0.0973 Range of lambda10: -36.8975 to 4.0556 lambda10 < L10: 0.99514 lambda10 >= L10: 0.00486
Report for threshold \(L_{10}=0.0973 \rightarrow E[v]=\)$-144.02
Just as in the single-test case, by changing the test threshold in the case of using both tests \( T_1+T_2 \), we’ve traded a higher confidence in using the test results for patients who are EP-positive (Ab = false negative rate decreases from about 1.75% → 0.39%) for a lower confidence in the results for patients who are EP-negative (Ba = false positive rate increases from about 0.16% → 0.49%). This and the increased cost assessment for false negatives means that the expected cost increases from \$119.11 to \$144.02 — which is still much better than the expected value from the one-test cost of \$813.91.
In numeric terms, for every 10 million patients, with 400 expected EP-positive patients, we can expect
- 393 will be correctly diagnosed as EP-positive, and 7 will be misdiagnosed as EP-negative, with \( L_{10} = 1.1013 \)
- 398 will be correctly diagnosed as EP-positive, and 2 will be misdiagnosed as EP-negative, with \( L_{10} = 0.0973 \)
(The cynical reader may conclude that, since the addition of test \( T_2 \) results in a \$670 decrease in expected cost over all potential patients, then the hospital should be charging \$770 for test \( T_2 \), not \$100.)
It’s worth noting again that we can never have perfect tests; there is always some chance of the test being incorrect. The only way to eliminate all false negatives is to increase the chances of false positives to 1. Choice of thresholds is always a compromise.
Another thing to remember is that real situations are seldom characterized perfectly by normal (Gaussian) distributions; the probability of events way out in the tails is usually higher than Gaussian because of black-swan events.
Remember: Por qué no los tres?
We’re almost done. We’ve just shown that by having everyone take both tests, \( T_1 \) and \( T_2 \), we can maximize expected value (minimize expected cost) over all patients.
But that wasn’t the original idea. Originally we were going to do this:
- Have everyone take the \$1 test \( T_1 \), which results in a measurement \( x \)
- If \( x \ge x_{H} \), diagnose as EP-positive, with no need for test \( T_2 \)
- If \( x \le x_{L} \), diagnose as EP-negative, with no need for test \( T_2 \)
- If \( x_{L} < x < x_{H} \), we will have the patient take the \$100 test \( T_2 \), which results in a measurement \( y \)
- If \( \lambda_{10}(x,y) \ge L_{10} \), diagnose as EP-positive
- If \( \lambda_{10}(x,y) < L_{10} \), diagnose as EP-negative
where \( \lambda_{10}(x,y) = \lambda(x,y) / \ln 10 = \log_{10} \frac{p_A(x,y)}{p_B(x,y)}. \)
Now we just need to calculate thresholds \( x_H \) and \( x_L \). These are going to need to have very low false positive and false negative rates, and they’re there to catch the “obvious” cases without the need to charge for (and wait for) test \( T_2 \).
We have the same kind of calculation as before. Let’s consider \( x=x_H \). It should be chosen so that with the correct threshold, there’s no change in expected value if we change the threshold by a small amount. (\( \partial E[v] / \partial x_H = 0 \)). We can do this by finding \( x_H \) such that, within a narrow range \( x_H < x < x_H+\Delta x \), the expected value is equal for both alternatives, namely whether we have them take test \( T_2 \), or diagnose them as EP-positive without taking test \( T_2 \).
Essentially we are determining thresholds \( x_H \) and \( x_L \) that, at each threshold, make the additional value of information gained from test \( T_2 \) (as measured by a change in expected value) equal to the cost of test \( T_2 \).
Remember that the total probability of \( x_H < x < x_H+\Delta x \) is \( (P_A p_A(x_H) + (1-P_A)p_B(x_H))\Delta x \), broken down into
- \( P_A p_A(x_H) \Delta x \) for EP-positive patients (A)
- \( (1-P_A) p_B(x_H) \Delta x \) for EP-negative patients (B)
Expected value \( V_1 \) for diagnosing as EP-positive (a) without taking test \( T_2 \), divided by \( \Delta x \) so we don’t have to keep writing it:
$$V_1(x_H) = P_A v_{Aa}p_A(x_H) + (1-P_A) v_{Ba} p_B(x_H)$$
where \( p_A(x), p_B(x) \) are Gaussian PDFs of the results of test \( T_1 \).
Expected value \( V_2 \) taking test \( T_2 \), which has value \( v(T_2)=-\$100 \):
$$\begin{aligned} V_2(x_0) &= v(T_2) + P_A E[v\ |\ x_0, A]p_A(x_0) + P_B E[v\ |\ x_0, B]p_B(x_0) \cr &= v(T_2) + P_A \left(v_{Aa}P_{2a}(x_0\ |\ A) + v_{Ab}P_{2b}(x_0\ |\ A)\right) p_A(x_0) \cr &+ (1-P_A)\left(v_{Ba} P_{2a}(x_0\ |\ B) + v_{Bb}P_{2b}(x_0\ | B)\right)p_B(x_0) \end{aligned}$$
where \( P_{2A}(x_0), P_{2B}(x_0) \) are the conditional probabilities of declaring the patient as EP-positive or EP-negative based on tests \( T_1, T_2 \), given that \( x=x_0 \). This happens to be the numbers we used for numerical integration in the previous section (where we were making all patients take tests \( T_1,T_2 \)).
When we have an optimal choice of threshold \( x_H \), the expected values will be equal: \( V_1(x_H) = V_2(x_H) \), because there is no advantage. If \( V_1(x_H) < V_2(x_H) \), then we haven’t chosen a good threshold, and \( x_H \) should be lower; if \( V_1(x_H) > V_2(x_H) \) then \( x_H \) should be higher. (Example: suppose that \( x_H = 55, V_1(x_H) = -400, \) and \( V_2(x_H) = -250. \) The way we’ve defined \( x_H \) is that any result of test \( T_1 \) where \( x > x_H \), the value of \( x \) is high enough that we’re better off — in other words, the expected value is supposed to be higher — just declaring diagnosis \( a \) instead of spending the extra \$100 to get a result from test \( T_2 \), given its expected value. In mathematical terms, \( V_1(x) > V_2(x) \) as long as \( x > x_H \). But we can choose \( x = x_H + \delta \) with arbitrarily small \( \delta \), and in this case we have \( V_1(x_H + \delta) > V_2(x_H+\delta) \) which contradicts \( V_1(x_H) < V_2(x_H) \). Nitpicky mathematicians: this means that the expected values \( V_1(x_H) \) and \( V_2(x_H) \) as functions of threshold \( x_H \) are continuous there. The proof should be a trivial exercise for the reader, right?)
So we just need to find the right value of \( x_H \) such that \( V_1(x_H) = V_2(x_H) \).
Presumably there is a way to determine a closed-form solution here, but I won’t even bother; we’ll just evaluate it numerically.
We’ll also cover the case where we need to find \( x_L \), where we decide just to make a diagnosis \( b \) if \( x < x_L \) based on \( T_1 \) alone, and the resulting expected value is
$$V_1(x_L) = P_A v_{Ab}p_A(x_L) + (1-P_A) v_{Bb} p_B(x_L),$$
otherwise if \( x \ge x_L \), pay the \$100 to take test \( T_2 \) with expected value \( V_2(x_L) \).
Then we just need to find \( x_L \) such that \( V_1(x_L) = V_2(x_L). \)
Let’s get an idea of what these functions look like for our example.
def compute_value_densities(which_one, distA, distB, Pa_total, value_matrix, vT2, L): """ Returns a function to compute value densities V0, V1 """ fxA = distA.slicefunc('x') fxB = distB.slicefunc('x') vAa = value_matrix[1,1] vBa = value_matrix[0,1] vAb = value_matrix[1,0] vBb = value_matrix[0,0] if which_one == 'H': vA1 = vAa vB1 = vBa elif which_one == 'L': vA1 = vAb vB1 = vBb else: raise ValueError("which_one must be 'H' or 'L'") normcdf = scipy.stats.norm.cdf C = distA.logpdf_coefficients - distB.logpdf_coefficients Q = Quadratic2D(*C) Qliss = Q.lissajous(L) xqr = Qliss.Rx xqmu = Qliss.x0 # Find the center and radius of the contour lambda(x,y)=L def v1v2(x_0, verbose=False): wA, muA, sigmaA = fxA(x_0) wB, muB, sigmaB = fxB(x_0) # wA = probability density at x = x_0 given case A # wB = probability density at x = x_0 given case B # muA, sigmaA describe the pdf p(y | A,x=x0) # muB, sigmaB describe the pdf p(y | B,x=x0) if x_0 < xqmu - xqr or x_0 > xqmu + xqr: # x is extreme enough that we always diagnose as "b" P2Aa = 0 P2Ab = 1 P2Ba = 0 P2Bb = 1 else: # Here we need to find the y-value thresholds theta = np.arccos((x_0-xqmu)/xqr) x1,y1 = Qliss(theta) x2,y2 = Qliss(-theta) assert np.abs(x_0-x1) < 1e-10*xqr, (x_0,x1,x2) assert np.abs(x_0-x2) < 1e-10*xqr, (x_0,x1,x2) if y1 > y2: y1,y2 = y2,y1 assert np.abs(Q(x_0,y1) - L) < 1e-10 assert np.abs(Q(x_0,y2) - L) < 1e-10 # Diagnose as "a" if between the thresholds, otherwise "b" P2Aa = normcdf(y2, muA, sigmaA) - normcdf(y1, muA, sigmaA) P2Ab = 1-P2Aa P2Ba = normcdf(y2, muB, sigmaB) - normcdf(y1, muB, sigmaB) P2Bb = 1-P2Ba # expected value given the patient has results of both tests # over the full range of y vA2 = vAa*P2Aa+vAb*P2Ab # given A, x_0 vB2 = vBa*P2Ba+vBb*P2Bb # given B, x_0 # Bayes' rule for conditional probability of A and B given x_0 PA = (Pa_total*wA)/(Pa_total*wA + (1-Pa_total)*wB) PB = 1-PA v1 = PA*vA1 + PB*vB1 v2 = vT2 + PA*vA2 + PB*vB2 if verbose: print which_one, x_0 print "PA|x0=",PA print vAa,vAb,vBa,vBb print P2Aa, P2Ab, P2Ba, P2Bb print "T1 only", vA1,vB1 print "T1+T2 ", vA2,vB2 print "v1=%g v2=%g v2-v1=%g" % (v1,v2,v2-v1) return v1,v2 return v1v2 Pa_total = 40e-6 value_matrix_T1 = np.array([[0,-5000],[-1e7, -1e5]]) vT2 = -100 L = compute_optimal_L(value_matrix_T1 + vT2, Pa_total) distA2 = Gaussian2D(mu_x=63,mu_y=57,sigma_x=5.3,sigma_y=4.1,rho=0.91,name='A (EP-positive)',id='A',color='red') distB2 = Gaussian2D(mu_x=48,mu_y=36,sigma_x=5.9,sigma_y=5.2,rho=0.84,name='B (EP-negative)',id='B',color='#8888ff') print "For L10=%.4f:" % (L/np.log(10)) for which in 'HL': print "" v1v2 = compute_value_densities(which, distA, distB, Pa_total, value_matrix_T1, vT2, L) for x_0 in np.arange(25,100.1,5): v1,v2 = v1v2(x_0) print "x_%s=%.1f, V1=%.4g, V2=%.4g, V2-V1=%.4g" % (which, x_0,v1,v2,v2-v1)
For L10=1.1013: x_H=25.0, V1=-5000, V2=-100, V2-V1=4900 x_H=30.0, V1=-5000, V2=-100, V2-V1=4900 x_H=35.0, V1=-5000, V2=-100.5, V2-V1=4900 x_H=40.0, V1=-5000, V2=-104.2, V2-V1=4896 x_H=45.0, V1=-5001, V2=-122.2, V2-V1=4879 x_H=50.0, V1=-5011, V2=-183.7, V2-V1=4828 x_H=55.0, V1=-5107, V2=-394.5, V2-V1=4713 x_H=60.0, V1=-5840, V2=-1354, V2-V1=4487 x_H=65.0, V1=-1.033e+04, V2=-6326, V2-V1=4008 x_H=70.0, V1=-2.878e+04, V2=-2.588e+04, V2-V1=2902 x_H=75.0, V1=-6.315e+04, V2=-6.185e+04, V2-V1=1301 x_H=80.0, V1=-8.695e+04, V2=-8.661e+04, V2-V1=339.4 x_H=85.0, V1=-9.569e+04, V2=-9.567e+04, V2-V1=26.9 x_H=90.0, V1=-9.843e+04, V2=-9.849e+04, V2-V1=-59.82 x_H=95.0, V1=-9.933e+04, V2=-9.942e+04, V2-V1=-85.24 x_H=100.0, V1=-9.967e+04, V2=-9.976e+04, V2-V1=-93.59 x_L=25.0, V1=-0.001244, V2=-100, V2-V1=-100 x_L=30.0, V1=-0.0276, V2=-100, V2-V1=-100 x_L=35.0, V1=-0.5158, V2=-100.5, V2-V1=-99.95 x_L=40.0, V1=-8.115, V2=-104.2, V2-V1=-96.09 x_L=45.0, V1=-107.5, V2=-122.2, V2-V1=-14.63 x_L=50.0, V1=-1200, V2=-183.7, V2-V1=1016 x_L=55.0, V1=-1.126e+04, V2=-394.5, V2-V1=1.087e+04 x_L=60.0, V1=-8.846e+04, V2=-1354, V2-V1=8.711e+04 x_L=65.0, V1=-5.615e+05, V2=-6326, V2-V1=5.551e+05 x_L=70.0, V1=-2.503e+06, V2=-2.588e+04, V2-V1=2.477e+06 x_L=75.0, V1=-6.121e+06, V2=-6.185e+04, V2-V1=6.059e+06 x_L=80.0, V1=-8.627e+06, V2=-8.661e+04, V2-V1=8.54e+06 x_L=85.0, V1=-9.547e+06, V2=-9.567e+04, V2-V1=9.451e+06 x_L=90.0, V1=-9.835e+06, V2=-9.849e+04, V2-V1=9.736e+06 x_L=95.0, V1=-9.93e+06, V2=-9.942e+04, V2-V1=9.83e+06 x_L=100.0, V1=-9.965e+06, V2=-9.976e+04, V2-V1=9.865e+06
import matplotlib.gridspec def fVdistort(V): return -np.log(-np.array(V+ofs)) yticks0 =np.array([1,2,5]) yticks = -np.hstack([yticks0 * 10**k for k in xrange(-1,7)]) for which in 'HL': fig = plt.figure(figsize=(6,6)) gs = matplotlib.gridspec.GridSpec(2,1,height_ratios=[2,1],hspace=0.1) ax=fig.add_subplot(gs[0]) x = np.arange(20,100,0.2) fv1v2 = compute_value_densities(which, distA, distB, Pa_total, value_matrix_T1, vT2, L) v1v2 = np.array([fv1v2(xi) for xi in x]) vlim = np.array([max(-1e6,v1v2.min()*1.01), min(-10,v1v2.max()*0.99)]) f = fVdistort ax.plot(x,f(v1v2[:,0]),x_H') if which == 'H' else ('"b"', '<x_L'))) ax.plot(x,f(v1v2[:,1]),label='T1+T2') ax.set_yticks(f(yticks)) ax.set_yticklabels('%.0f' % y for y in yticks) ax.set_ylim(f(vlim)) ax.set_ylabel('$E[v]$',fontsize=12) ax.grid(True) ax.legend(labelspacing=0, fontsize=10, loc='lower left' if which=='H' else 'upper right') [label.set_visible(False) for label in ax.get_xticklabels()] ax2=fig.add_subplot(gs[1], sharex=ax) ax2.plot(x,v1v2[:,1]-v1v2[:,0]) ax2.set_ylim(-500,2000) ax2.grid(True) ax2.set_ylabel('$\\Delta E[v]$') ax2.set_xlabel('$x_%s$' % which, fontsize=12)
It looks like for this case (\( L_{10}=1.1013 \)), we should choose \( x_L \approx 45 \) and \( x_H \approx 86 \).
These edge cases where \( x < x_L \) or \( x > x_H \) don’t save a lot of money, at best just the \$100 \( T_2 \) test cost… so a not-quite-as-optimal but simpler case would be to always run both tests. Still, there’s a big difference between going to the doctor and paying \$1 rather than \$101… whereas once you’ve paid a \$100,000 hospital bill, what’s an extra \$100 among friends?
Between those thresholds, where test \( T_1 \) is kind of unhelpful, the benefit of running both tests is enormous: for EP-positive patients we can help minimize those pesky false negatives, saving hospitals millions in malpractice charges and helping those who would otherwise have grieving families; for EP-negative patients we can limit the false positives and save them the \$5000 and anguish of a stressful hospital stay.
Putting it all together
Now we can show our complete test protocol on one graph and one chart. Below the colored highlights show four regions:
- blue: \( b_1 \) — EP-negative diagnosis after taking test \( T_1 \) only, with \( x<x_L \)
- green: \( b_2 \) — EP-negative diagnosis after taking tests \( T_1, T_2 \), with \( x_L \le x \le x_H \) and \( \lambda_{10} < L_{10} \)
- yellow: \( a_2 \) — EP-positive diagnosis after taking tests \( T_1, T_2 \), with \( x_L \le x \le x_H \) and \( \lambda_{10} \ge L_{10} \)
- red: \( a_1 \) — EP-positive diagnosis after taking tests \( T_1 \) only, with \( x > x_H \)
import matplotlib.patches as patches import matplotlib.path import scipy.optimize Path = matplotlib.path.Path def show_complete_chart(xydistA, xydistB, Q, L, Pa_total, value_matrix_T1, vT2, return_info=False): fig = plt.figure() ax = fig.add_subplot(1,1,1) xlim = (0,100) ylim = (0,100) separation_plot(xydistA, xydistB, Q, L, ax=ax, xlim=xlim, ylim=ylim) ax.set_xticks(np.arange(0,101.0,10)) ax.set_yticks(np.arange(0,101.0,10)) # Solve for xL and xH _,_,distA = xydistA _,_,distB = xydistB v1v2 = [compute_value_densities(which, distA, distB, Pa_total, value_matrix_T1, vT2, L) for which in 'LH'] def fdelta(f): def g(x): y1,y2 = f(x) return y1-y2 return g xL = scipy.optimize.brentq(fdelta(v1v2[0]), 0, 100) xH = scipy.optimize.brentq(fdelta(v1v2[1]), 0, 100) # Show the full matrix of possibilities: # compute 2x4 confusion matrix C = [] for dist in [distB, distA]: distx = dist.project('x') pa2,pb2 = estimate_separation_numerical(dist, Q, L, xL, xH, Nintervals=2500, return_pair=True) row = [distx.cdf(xL), pb2, pa2, 1-distx.cdf(xH)] C.append(row) # compute 2x4 value matrix V = value_matrix_T1.repeat(2,axis=1) V[:,1:3] += vT2 display(show_binary_matrix(confusion_matrix=C, threshold='x_L=%.3f, x_H=%.3f, L_{10}=%.3f' %(xL,xH,L/np.log(10)), distributions=[distB,distA], outcome_ids=['b1','b2','a2','a1'], ppos=Pa_total, vmatrix=V, special_format={'v':'$%.2f', 'J':['%.8f','%.8f','%.8f','%.3g']})) # Highlight each of the areas x0,x1 = xlim y0,y1 = ylim # area b1: x < xL rect = patches.Rectangle((x0,y0),xL-x0,y1-y0,linewidth=0,edgecolor=None,facecolor='#8888ff',alpha=0.1) ax.add_patch(rect) # area a1: x > xH rect = patches.Rectangle((xH,y0),x1-xH,y1-y0,linewidth=0,edgecolor=None,facecolor='red',alpha=0.1) ax.add_patch(rect) for x in [xL,xH]: ax.plot([x,x],[y0,y1],color='gray') # area a2: lambda(x,y) > L xc,yc = Q.contour(L) ii = (xc > xL-10) & (xc < xH + 10) xc = xc[ii] yc = yc[ii] xc = np.maximum(xc,xL) xc = np.minimum(xc,xH) poly2a = patches.Polygon(np.vstack([xc,yc]).T, edgecolor=None, facecolor='yellow',alpha=0.5) ax.add_patch(poly2a) # area b2: lambda(x,y) < L xy = [] op = [] i = 0 for x,y in zip(xc,yc): i += 1 xy.append((x,y)) op.append(Path.MOVETO if i == 1 else Path.LINETO) xy.append((0,0)) op.append(Path.CLOSEPOLY) xy += [(xL,y0),(xL,y1),(xH,y1),(xH,y0),(0,0)] op += [Path.MOVETO, Path.LINETO, Path.LINETO, Path.LINETO, Path.CLOSEPOLY] patch2b = patches.PathPatch(Path(xy,op), edgecolor=None, facecolor='#00ff00',alpha=0.1) ax.add_patch(patch2b) # add labels style = {'fontsize': 28, 'ha':'center'} ax.text((x0+xL)/2,y0*0.3+y1*0.7,'$b_1$', **style) ax.text(xc.mean(), yc.mean(), '$a_2$', **style) a = 0.3 if yc.mean() > (x0+x1)/2 else 0.7 yb2 = a*y1 + (1-a)*y0 ax.text(xc.mean(), yb2, '$b_2$',**style) xa1 = (xH + x1) / 2 ya1 = max(30, min(90, Q.constrain(x=xa1).x0)) ax.text(xa1,ya1,'$a_1$',**style) if return_info: C = np.array(C) J = C * [[1-Pa_total],[Pa_total]] return dict(C=C,J=J,V=V,xL=xL,xH=xH,L=L) value_matrix_T1 = np.array([[0,-5000],[-1e7, -1e5]]) value_matrix_T2 = value_matrix_T1 - 100 L = compute_optimal_L(value_matrix_T2, Pa_total) info37 = show_complete_chart((xA,yA,distA),(xB,yB,distB), Q, L, Pa_total, value_matrix_T1, vT2, return_info=True)
Report for threshold \(x_L=45.287, x_H=85.992, L_{10}=1.101 \rightarrow E[v]=\)$-46.88
We can provide the same information but with the false negative cost (Ab = wrongly diagnosed EP-negative) at \$100 million:
value_matrix_T1 = np.array([[0,-5000],[-1e8, -1e5]]) value_matrix_T2 = value_matrix_T1 - 100 L = compute_optimal_L(value_matrix_T2, Pa_total) info38 = show_complete_chart((xA,yA,distA),(xB,yB,distB), Q, L, Pa_total, value_matrix_T1, vT2, return_info=True)
Report for threshold \(x_L=40.749, x_H=85.108, L_{10}=0.097 \rightarrow E[v]=\)$-99.60
Do we need test \( T_1 \)?
If adding test \( T_2 \) is so much better than test \( T_1 \) alone, why do we need test \( T_1 \) at all? What if we just gave everyone test \( T_2 \)?
y = np.arange(0,100,0.1) distA_T2 = distA.project('y') distB_T2 = distB.project('y') show_binary_pdf(distA_T2, distB_T2, y, xlabel = '$T_2$ test result $y$') for false_negative_cost in [10e6, 100e6]: value_matrix_T2 = np.array([[0,-5000],[-false_negative_cost, -1e5]]) - 100 threshold_info = find_threshold(distB_T2, distA_T2, Pa_total, value_matrix_T2) y0 = [yval for yval,_ in threshold_info if yval > 20 and yval < 80][0] C = analyze_binary(distB_T2, distA_T2, y0) print "False negative cost = $%.0fM" % (false_negative_cost/1e6) display(show_binary_matrix(C, 'y_0=%.2f' % y0, [distB_T2, distA_T2], 'ba', Pa_total, value_matrix_T2, special_format={'v':'$%.2f'})) info = gather_info(C,value_matrix_T2,Pa_total,y0=y0) if false_negative_cost == 10e6: info47 = info else: info48 = info
False negative cost = $10M
Report for threshold \(y_0=50.14 \rightarrow E[v]=\)$-139.03
False negative cost = $100M
Report for threshold \(y_0=47.73 \rightarrow E[v]=\)$-211.68
Hmm. That seems better than the \( T_1 \) test… but the confusion matrix doesn’t seem as good as using both tests.
d prime \( (d’) \): Are two tests are always better than one?
Why would determining a diagnosis based on both tests \( T_1 \) and \( T_2 \) be better than from either test alone?
Let’s graph the PDFs of three different measurements:
- \( x \) (the result of test \( T_1 \))
- \( y \) (the result of test \( T_2 \))
- \( u = -0.88x + 1.88y \) (a linear combination of the two measurements)
We’ll also calculate a metric for each of the three measurements,
$$d’ = \frac{\mu_A - \mu_B}{\sqrt{\frac{1}{2}(\sigma_A{}^2 + \sigma_B{}^2)}}$$
which is a measure of the “distinguishability” between two populations which each have normal distributions. Essentially it is a unitless “separation coefficient” that measures the distances between the means as a multiple of the “average” standard deviation. It is also invariant under scaling and translation — if we figure out the value of \( d’ \) for some measurement \( x \), then any derived measurement \( u = ax + c \) has the same value for \( d’ \) as long as \( a \ne 0 \).
(We haven’t talked about derived measurements like \( u \), but for a 2-D Gaussian distribution, if \( u=ax+by+c \) then:
$$\begin{aligned} \mu_u &= E[u] = aE[x]+bE[y] = a\mu_x + b\mu_y + c\cr \sigma_u{}^2 &= E[(u-\mu_u)^2] \cr &= E[a^2(x-\mu_x)^2 + 2ab(x-\mu_x)(y-\mu_y) + b^2(y-\mu_y)^2] \cr &= a^2E[(x-\mu_x)^2] + 2abE[(x-\mu_x)(y-\mu_y)] + b^2E[(y-\mu_y)^2] \cr &= a^2 \sigma_x{}^2 + 2ab\rho\sigma_x\sigma_y + b^2\sigma_y{}^2 \end{aligned}$$
or, alternatively using matrix notation, \( \sigma_u{}^2 = \mathrm{v}^T \mathrm{S} \mathrm{v} \) where \( \mathrm{v} = \begin{bmatrix}a& b\end{bmatrix}^T \) is the vector of weighting coefficients, and \( \mathrm{S} = \operatorname{cov}(x,y) = \begin{bmatrix}\sigma_x{}^2 & \rho\sigma_x\sigma_y \cr \rho\sigma_x\sigma_y & \sigma_y{}^2\end{bmatrix} \). Yep, there’s the covariance matrix again.)
def calc_dprime(distA,distB,projection): """ calculates dprime for two distributions, given a projection P = [cx,cy] that defines u = cx*x + cy*y """ distAp = distA.project(projection) distBp = distB.project(projection) return (distAp.mu - distBp.mu)/np.sqrt( (distAp.sigma**2 + distBp.sigma**2)/2.0 ) def show_dprime(distA, distB, a,b): print "u=%.6fx%+.6fy" % (a,b) x = np.arange(0,100,0.1) fig = plt.figure() ax = fig.add_subplot(1,1,1) for projname, projection, linestyle in [ ('x','x',':'), ('y','y','--'), ('u',[a,b],'-')]: distAp = distA.project(projection) distBp = distB.project(projection) dprime = calc_dprime(distA, distB, projection) print "dprime(%s)=%.4f" % (projname,dprime) for dist in [distAp,distBp]: ax.plot(x,dist.pdf(x), color=dist.color, linestyle=linestyle, label='$%s$: %s, $\\mu=$%.1f, $\\sigma=$%.2f' % (projname, dist.id, dist.mu, dist.sigma)) ax.set_ylabel('probability density') ax.set_xlabel('measurement $(x,y,u)$') ax.set_ylim(0,0.15) ax.legend(labelspacing=0, fontsize=10,loc='upper left'); show_dprime(distA, distB, -0.88,1.88)
u=-0.880000x+1.880000y dprime(x)=2.6747 dprime(y)=4.4849 dprime(u)=5.1055
We can get a better separation between alternatives, as measured by \( d’ \), through this linear combination of \( x \) and \( y \) than by either one alone. What’s going on, and where did the equation \( u=-0.88x + 1.88y \) come from?
Can we do better than that? How about \( u=-0.98x + 1.98y \)? or \( u=-0.78x + 1.78y \)?
show_dprime(distA, distB, -0.98,1.98)
u=-0.980000x+1.980000y dprime(x)=2.6747 dprime(y)=4.4849 dprime(u)=5.0997
show_dprime(distA, distB, -0.78,1.78)
u=-0.780000x+1.780000y dprime(x)=2.6747 dprime(y)=4.4849 dprime(u)=5.0988
Hmm. These aren’t quite as good; the value for \( d’ \) is lower. How do we know how to maximize \( d’ \)?
Begin grungy algebra
There doesn’t look like a simple intuitive way to find the best projection. At first I thought of modeling this as \( v = x \cos\theta + y \sin\theta \) for some angle \( \theta \). But this didn’t produce any easy answer. I muddled my way to an answer by looking for patterns that helped to cancel out some of the grunginess.
One way to maximize \( d’ \) is to write it as a function of \( \theta \) with \( v=ax+by, a = a_1\cos \theta+a_2\sin\theta, b = b_1\cos\theta+b_2\sin\theta \) for some convenient \( a_1, a_2, b_1, b_2 \) to make the math nice. We’re going to figure out \( d’ \) as a function of \( a,b \) in general first:
$$\begin{aligned} d’ &= \frac{\mu_{vA} - \mu_{vB}}{\sqrt{\frac{1}{2}(\sigma_{vA}{}^2 + \sigma_{vB}{}^2)}} \cr &= \frac{a(\mu_{xA}-\mu_{xB})+b(\mu_{yA} - \mu_{yB})}{\sqrt{\frac{1}{2}(a^2(\sigma_{xA}^2 + \sigma_{xB}^2) + 2ab(\rho_A\sigma_{xA}\sigma_{yA} + \rho_B\sigma_{xB}\sigma_{yB}) + b^2(\sigma_{yA}^2+\sigma_{yB}^2))}} \cr &= \frac{a(\mu_{xA}-\mu_{xB})+b(\mu_{yA} - \mu_{yB})}{\sqrt{\frac{1}{2}f(a,b)}} \cr \end{aligned}$$
with \( f(a,b) = a^2(\sigma_{xA}^2 + \sigma_{xB}^2) + 2ab(\rho_A\sigma_{xA}\sigma_{yA} + \rho_B\sigma_{xB}\sigma_{yB}) + b^2(\sigma_{yA}^2+\sigma_{yB}^2). \)
Yuck.
Anyway, if we can make the denominator constant as a function of \( \theta \), then the numerator is easy to maximize.
If we define
$$\begin{aligned} K_x &= \sqrt{\sigma_{xA}^2 + \sigma_{xB}^2} \cr K_y &= \sqrt{\sigma_{yA}^2 + \sigma_{yB}^2} \cr R &= \frac{\rho_A\sigma_{xA}\sigma_{yA} + \rho_B\sigma_{xB}\sigma_{yB}}{K_xK_y} \end{aligned}$$
then \( f(a,b) = a^2K_x{}^2 + 2abRK_xK_y + b^2K_y{}^2 \) which is easier to write than having to carry around all those \( \mu \) and \( \sigma \) terms.
If we let \( a = \frac{1}{\sqrt{2}K_x}(c \cos \theta + s \sin\theta) \) and \( b = \frac{1}{\sqrt{2}K_y}(c \cos \theta - s\sin\theta) \) then
$$\begin{aligned} f(a,b) &= \frac{1}{2}(c^2 \cos^2\theta + s^2\sin^2\theta + 2cs\cos\theta\sin\theta) \cr &+ \frac{1}{2}(c^2 \cos^2\theta + s^2\sin^2\theta - 2cs\cos\theta\sin\theta) \cr &+R(c^2 \cos^2\theta - s^2\sin^2\theta) \cr &= (1+R)c^2 \cos^2\theta + (1-R)s^2\sin^2\theta \end{aligned}$$
which is independent of \( \theta \) if \( (1+R)c^2 = (1-R)s^2 \). If we let \( c = \cos \phi \) and \( s=\sin \phi \) then some nice things all fall out:
- \( f(a,b) \) is constant if \( \tan^2\phi = \frac{1+R}{1-R} \), in other words \( \phi = \tan^{-1} \sqrt{\frac{1+R}{1-R}} \)
- \( c = \cos\phi = \sqrt{\frac{1-R}{2}} \) (hint: use the identities \( \sec^2 \theta = \tan^2 \theta + 1 \) and \( \cos \theta = 1/\sec \theta \))
- \( s = \sin\phi = \sqrt{\frac{1+R}{2}} \)
- \( f(a,b) = (1-R^2)/2 \)
- \( a = \frac{1}{\sqrt{2}K_x}\cos (\theta - \phi) = \frac{1}{\sqrt{2}K_x}(\cos \phi \cos \theta + \sin\phi \sin\theta)= \frac{1}{2K_x}(\sqrt{1-R} \cos \theta + \sqrt{1+R} \sin\theta) \)
- \( b = \frac{1}{\sqrt{2}K_y}\cos (\theta + \phi) = \frac{1}{\sqrt{2}K_y}(\cos \phi \cos \theta - \sin\phi \sin\theta)= \frac{1}{2K_y}(\sqrt{1-R} \cos \theta - \sqrt{1+R} \sin\theta) \)
and when all is said and done we get
$$\begin{aligned} d’ &= \frac{a(\mu_{xA}-\mu_{xB})+b(\mu_{yA} - \mu_{yB})}{\sqrt{\frac{1}{2}f(a,b)}} \cr &= \frac{a(\mu_{xA}-\mu_{xB})+b(\mu_{yA} - \mu_{yB})}{\frac{1}{2}\sqrt{1-R^2}} \cr &= K_c \cos\theta + K_s \sin\theta \end{aligned}$$
if
$$\begin{aligned} \delta_x &= \frac{\mu_{xA}-\mu_{xB}}{K_x} = \frac{\mu_{xA}-\mu_{xB}}{\sqrt{\sigma_{xA}^2 + \sigma_{xB}^2}} \cr \delta_y &= \frac{\mu_{yA} - \mu_{yB}}{K_y} = \frac{\mu_{yA} - \mu_{yB}}{\sqrt{\sigma_{yA}^2 + \sigma_{yB}^2}} \cr K_c &= \sqrt{\frac{1-R}{1-R^2}} \left(\delta_x + \delta_y \right) \cr &= \frac{1}{\sqrt{1+R}} \left(\delta_x + \delta_y \right) \cr K_s &= \sqrt{\frac{1+R}{1-R^2}} \left(\delta_x - \delta_y \right) \cr &= \frac{1}{\sqrt{1-R}} \left(\delta_x - \delta_y \right) \cr \end{aligned}$$
Then \( d’ \) has a maximum value of \( D = \sqrt{K_c{}^2 + K_s{}^2} \) at \( \theta = \tan^{-1}\dfrac{K_s}{K_c} \), where \( \cos \theta = \dfrac{K_c}{\sqrt{K_s^2 + K_c^2}} \) and \( \sin \theta = \dfrac{K_s}{\sqrt{K_s^2 + K_c^2}}. \)
Plugging into \( a \) and \( b \) we get
$$\begin{aligned} a &= \frac{1}{2K_x}(\sqrt{1-R} \cos \theta + \sqrt{1+R} \sin\theta)\cr &= \frac{1}{2K_x}\left(\frac{(1-R)(\delta_x+\delta_y)}{\sqrt{1-R^2}}+\frac{(1+R)(\delta_x-\delta_y)}{\sqrt{1-R^2}}\right)\cdot\frac{1}{\sqrt{K_s^2 + K_c^2}} \cr &= \frac{\delta_x - R\delta_y}{K_xD\sqrt{1-R^2}}\cr b &= \frac{1}{2K_y}(\sqrt{1-R} \cos \theta - \sqrt{1+R} \sin\theta)\cr &= \frac{1}{2K_y}\left(\frac{(1-R)(\delta_x+\delta_y)}{\sqrt{1-R^2}}-\frac{(1+R)(\delta_x-\delta_y)}{\sqrt{1-R^2}}\right)\cdot\frac{1}{\sqrt{K_s^2 + K_c^2}} \cr &= \frac{-R\delta_x + \delta_y}{K_yD\sqrt{1-R^2}} \end{aligned}$$
We can also solve \( D \) in terms of \( \delta_x, \delta_y, \) and \( R \) to get
$$\begin{aligned} D &= \frac{1}{\sqrt{1-R^2}}\sqrt{(1-R)(\delta_x+\delta_y)^2 + (1+R)(\delta_x-\delta_y)^2} \cr &= }} \end{aligned}$$
Let’s try it!
Kx = np.hypot(distA.sigma_x,distB.sigma_x) Ky = np.hypot(distA.sigma_y,distB.sigma_y) R = (distA.rho*distA.sigma_x*distA.sigma_y + distB.rho*distB.sigma_x*distB.sigma_y)/Kx/Ky Kx, Ky, R
(7.9309520235593407, 6.6219332524573211, 0.86723211589363869)
dmux = (distA.mu_x - distB.mu_x)/Kx dmuy = (distA.mu_y - distB.mu_y)/Ky Kc = (dmux + dmuy)/np.sqrt(1+R) Ks = (dmux - dmuy)/np.sqrt(1-R) Kc,Ks
(3.7048851247525367, -3.5127584699841927)
theta = np.arctan(Ks/Kc) a = 1.0/2/Kx*(np.sqrt(1-R)*np.cos(theta) + np.sqrt(1+R)*np.sin(theta) ) b = 1.0/2/Ky*(np.sqrt(1-R)*np.cos(theta) - np.sqrt(1+R)*np.sin(theta) ) dprime = np.hypot(Kc,Ks) dprime2 = Kc*np.cos(theta) + Ks*np.sin(theta) assert abs(dprime - dprime2) < 1e-7 delta_x = (distA.mu_x - distB.mu_x)/np.hypot(distA.sigma_x,distB.sigma_x) delta_y = (distA.mu_y - distB.mu_y)/np.hypot(distA.sigma_y,distB.sigma_y) dd = np.sqrt(2*delta_x**2 - 4*R*delta_x*delta_y + 2*delta_y**2) dprime3 = dd/np.sqrt(1-R*R) assert abs(dprime - dprime3) < 1e-7 a2 = (delta_x - R*delta_y)/Kx/dd b2 = (-R*delta_x + delta_y)/Ky/dd assert abs(a-a2) < 1e-7 assert abs(b-b2) < 1e-7 print "theta=%.6f a=%.6f b=%.6f d'=%.6f" % (theta, a, b, dprime) # scale a,b such that their sum is 1.0 print " also a=%.6f b=%.6f is a solution with a+b=1" % (a/(a+b), b/(a+b))
theta=-0.758785 a=-0.042603 b=0.090955 d'=5.105453 also a=-0.881106 b=1.881106 is a solution with a+b=1
And out pops \( u\approx -0.88x + 1.88y \).
End grungy algebra
What if we use \( u=-0.8811x + 1.8811y \) as a way to combine the results of tests \( T_1 \) and \( T_2 \)? Here is a superposition of lines with constant \( u \):
# Lines of constant u fig = plt.figure() ax = fig.add_subplot(1,1,1) separation_plot((xA,yA,distA),(xB,yB,distB), Q, L, ax=ax) xmax = ax.get_xlim()[1] ymax = ax.get_ylim()[1] ua = a/(a+b) ub = b/(a+b) for u in np.arange(ua*xmax,ub*ymax+0.01,10): x1 = min(xmax,(u-ub*ymax)/ua) x0 = max(0,u/ua) x = np.array([x0,x1]) ax.plot(x,(u-ua*x)/ub,color='orange')
To summarize:
$$\begin{aligned} R &= \frac{\rho_A\sigma_{xA}\sigma_{yA} + \rho_B\sigma_{xB}\sigma_{yB}}{\sqrt{(\sigma_{xA}^2 + \sigma_{xB}^2)(\sigma_{yA}^2 + \sigma_{yB}^2)}} \cr \delta_x &= \frac{\mu_{xA}-\mu_{xB}}{\sqrt{\sigma_{xA}^2 + \sigma_{xB}^2}} \cr \delta_y &= \frac{\mu_{yA} - \mu_{yB}}{\sqrt{\sigma_{yA}^2 + \sigma_{yB}^2}} \cr \max d’ = D &= }}\cr \end{aligned}$$
and we can calculate a new derived measurement \( u=ax+by \) which has separation coefficient \( d’ \) between its distributions under the conditions \( A \) and \( B \).
u = np.arange(0,100,0.1) projAB = [-0.8811, 1.8811] distA_T1T2lin = distA.project(projAB) distB_T1T2lin = distB.project(projAB) show_binary_pdf(distA_T1T2lin, distB_T1T2lin, y, xlabel = '$u = %.4fx + %.4fy$' % tuple(projAB)) for false_negative_cost in [10e6, 100e6]: value_matrix_T2 = np.array([[0,-5000],[-false_negative_cost, -1e5]]) - 100 threshold_info = find_threshold(distB_T1T2lin, distA_T1T2lin, Pa_total, value_matrix_T2) u0 = [uval for uval,_ in threshold_info if uval > 20 and uval < 80][0] C = analyze_binary(distB_T1T2lin, distA_T1T2lin, u0) print "False negative cost = $%.0fM" % (false_negative_cost/1e6) display(show_binary_matrix(C, 'u_0=ax+by=%.2f' % u0, [distB_T1T2lin, distA_T1T2lin], 'ba', Pa_total, value_matrix_T2, special_format={'v':'$%.2f'})) info = gather_info(C,value_matrix_T2,Pa_total,u0=u0) if false_negative_cost == 10e6: info57 = info else: info58 = info
False negative cost = $10M
Report for threshold \(u_0=ax+by=50.42 \rightarrow E[v]=\)$-119.26
False negative cost = $100M
Report for threshold \(u_0=ax+by=48.22 \rightarrow E[v]=\)$-144.54
One final subtlety
Before we wrap up the math, there’s one more thing that is worth a brief mention.
When we use both tests with the quadratic function \( \lambda(x,y) \), there’s a kind of funny region we haven’t talked about. Look on the graph below, at point \( P = (x=40, y=70) \):
fig = plt.figure() ax = fig.add_subplot(1,1,1) separation_plot((xA,yA,distA),(xB,yB,distB), Q, L, ax=ax) P=Coordinate2D(40,70) ax.plot(P.x,P.y,'.',color='orange') Ptext = Coordinate2D(P.x-10,P.y) ax.text(Ptext.x,Ptext.y,'$P$',fontsize=20, ha='right',va='center') ax.annotate('',xy=P,xytext=Ptext, arrowprops=dict(facecolor='black', width=1, headwidth=5, headlength=8, shrink=0.05));
This point is outside the contour \( \lambda_{10} > L_{10} \), so that tells us we should give a diagnosis of EP-negative. But this point is closer to the probability “cloud” of EP-positive. Why?
for dist in [distA, distB]: print("Probability density at P of %s = %.3g" % (dist.name, dist.pdf(P.x,P.y))) print(dist)
Probability density at P of A (EP-positive) = 6.28e-46 Gaussian(mu_x=55, mu_y=57, sigma_x=5.3, sigma_y=4.1, rho=0.91, name='A (EP-positive)', id='A', color='red') Probability density at P of B (EP-negative) = 2.8e-34 Gaussian(mu_x=40, mu_y=36, sigma_x=5.9, sigma_y=5.2, rho=0.84, name='B (EP-negative)', id='B', color='#8888ff')
The probability density is smaller at P for the EP-positive case, even though this point is closer to the corresponding probability distribution. This is because the distribution is narrower; \( \sigma_x \) and \( \sigma_y \) are both smaller for the EP-positive case.
As a thought experiment, imagine the following distributions A and B, where B is very wide (\( \sigma=5 \)) compared to A (\( \sigma = 0.5 \)):
dist_wide = Gaussian1D(20,5,'B (wide)',id='B',color='green') dist_narrow = Gaussian1D(25,0.5,'A (narrow)',id='A',color='red') x = np.arange(0,40,0.1) show_binary_pdf(dist_wide, dist_narrow, x) plt.xlabel('x');
In this case, if we take some sample measurement \( x \), a classification of \( A \) makes sense only if the measured value \( x \) is near \( A \)’s mean value \( \mu=25 \), say from 24 to 26. Outside that range, a classification of \( B \) makes sense, not because the measurement is closer to the mean of \( B \), but because the expected probability of \( B \) given \( x \) is greater. This holds true even for values within a few standard deviations from the mean, like \( x=28 = \mu_B + 1.6\sigma_B \), because the distribution of \( A \) is so narrow.
We’ve been proposing tests where there is a single threshold — for example diagnose as \( a \) if \( x > x_0 \), otherwise diagnose as \( b \) — but there are fairly simple cases where two thresholds are required. In fact, this is true in general when the standard deviations are unequal; if you get far enough away from the means, the probability of the wider distribution is greater.
You might think, well, that’s kind of strange, if I’m using the same piece of equipment to measure all the sample data, why would the standard deviations be different? A digital multimeter measuring voltage, for example, has some inherent uncertainty. But sometimes the measurement variation is due to differences in the populations themselves. For example, consider the wingspans of two populations of birds, where \( A \) consists of birds that are pigeons and \( B \) consists of birds that are not pigeons. The \( B \) population will have a wider range of measurements simply because this population is more heterogeneous.
Please note, by the way, that the Python functions I wrote to analyze the bivariate Gaussian situation make the assumption that the standard deviations are unequal and the log-likelihood ratio \( \lambda(x,y) \) is either concave upwards or concave downwards — in other words, the \( A \) distribution has lower values of both \( \sigma_x \) and \( \sigma_y \) than the \( B \) distribution, or it has higher values of both \( \sigma_x \) and \( \sigma_y \). In these cases, the contours of constant \( \lambda \) are ellipses. In general, they may be some other conic section (lines, parabolas, hyperbolas) but I didn’t feel like trying to achieve correctness in all cases, for the purposes of this article.
So which test is best?
Let’s summarize the tests we came up with. We have 10 different tests, namely each of the following with false-negative costs of \$10 million and \$100 million:
- test \( T_1 \) only
- tests \( T_1 \) and \( T_2 \) – quadratic function of \( x,y \)
- test \( T_1 \) first, then \( T_2 \) if needed
- test \( T_2 \) only
- tests \( T_1 \) and \( T_2 \) – linear function of \( x,y \)
import pandas as pd def present_info(info): C = info['C'] J = info['J'] V = info['V'] if 'x0' in info: threshold = '\\( x_0 = %.2f \\)' % info['x0'] t = 1 Bb1 = J[0,0] Bb2 = 0 elif 'xL' in info: threshold = ('\\( x_L = %.2f, x_H=%.2f, L_{10}=%.4f \\)' % (info['xL'],info['xH'],info['L']/np.log(10))) t = 3 Bb1 = J[0,0] Bb2 = J[0,1] elif 'L' in info: threshold = '\\( L_{10}=%.4f \\)' % (info['L']/np.log(10)) t = 2 Bb1 = 0 Bb2 = J[0,0] elif 'y0' in info: threshold = '\\( y_0 = %.2f \\)' % info['y0'] t = 4 Bb1 = 0 Bb2 = J[0,0] elif 'u0' in info: threshold = '\\( u_0 = ax+by = %.2f \\)' % info['u0'] t = 4 Bb1 = 0 Bb2 = J[0,0] nc = J.shape[1] return {'Exp. cost':'$%.2f' % (-J*V).sum(), 'Threshold': threshold, 'N(Ab)': int(np.round(10e6*J[1,:nc//2].sum())), 'N(Ba)': int(np.round(10e6*J[0,nc//2:].sum())), 'N(Aa)': int(np.round(10e6*J[1,nc//2:].sum())), 'P(b|A)': '%.2f%%' % (C[1,:nc//2].sum() * 100), 'P(Bb,$0)': '%.2f%%' % (Bb1*100), 'P(Bb,$100)': '%.2f%%' % (Bb2*100), } tests = ['\\(T_1\\)', '\\(T_1 + T_2\\)', '\\(T_1 \\rightarrow T_2 \\)', '\\(T_2\\)', '\\(T_1 + T_2\\) (linear)'] print "Exp. cost = expected cost above test T1 ($1)" print "N(Aa) = expected number of correctly-diagnosed EP-positives per 10M patients" print "N(Ab) = expected number of false negatives per 10M patients" print "N(Ba) = expected number of false positives per 10M patients" print("P(Bb,$0) = percentage of patients correctly diagnosed EP-negative,\n"+ " no additional cost after test T1") print("P(Bb,$100) = percentage of patients correctly diagnosed EP-negative,\n"+ " test T2 required") print "P(b|A) = conditional probability of a false negative for EP-positive patients" print "T1 -> T2: test T1 followed by test T2 if needed" df = pd.DataFrame([present_info(info) for info in [info17,info27,info37,info47,info57, info18,info28,info38,info48,info58]], index=['\\$%dM, %s' % (cost,test) for cost in [10,100] for test in tests]) colwidths = [12,10,7,7,10,10,10,10,24] colwidthstyles = [{'selector':'thead th.blank' if i==0 else 'th.col_heading.col%d' % (i-1), 'props':[('width','%d%%'%w)]} for i,w in enumerate(colwidths)] df.style.set_table_styles([{'selector':' ','props': {'width':'100%', 'table-layout':'fixed', }.items()}, {'selector':'thead th', 'props':[('font-size','90%')]}, {'selector': 'td span.MathJax_CHTML,td span.MathJax,th span.MathJax_CHTML,th span.MathJax', 'props':[('white-space','normal')]}, {'selector':'td.data.col7', 'props':[('font-size','90%')]}] + colwidthstyles)
Exp. cost = expected cost above test T1 ($1) N(Aa) = expected number of correctly-diagnosed EP-positives per 10M patients N(Ab) = expected number of false negatives per 10M patients N(Ba) = expected number of false positives per 10M patients P(Bb,$0) = percentage of patients correctly diagnosed EP-negative, no additional cost after test T1 P(Bb,$100) = percentage of patients correctly diagnosed EP-negative, test T2 required P(b|A) = conditional probability of a false negative for EP-positive patients T1 -> T2: test T1 followed by test T2 if needed
There! We can keep expected cost down by taking the \( T_1 \rightarrow T_2 \) approach (test 2 only if \( x_L \le x \le x_H \)).
With \$10M cost of a false negative, we can expect over 81% of patients will be diagnosed correctly as EP-negative with just the \$1 cost of the eyeball photo test, and only 18 patients out of ten million (4.5% of EP-positive patients) will experience a false negative.
With \$100M cost of a false negative, we can expect over 55% of patients will be diagnosed correctly as EP-negative with just the \$1 cost of the eyeball photo test, and only 3 patients out of ten million (0.7% of EP-positive patients) will experience a false negative.
And although test \( T_2 \) alone is better than test \( T_1 \) alone, both in keeping expected costs down, and in reducing the incidence of both false positives and false negatives, it’s pretty clear that relying on both tests is the most attractive option.
There is not too much difference between the combination of tests \( T_1+T_2 \) using \( \lambda_{10}(x,y) > L_{10} \) as a quadratic function of \( x \) and \( y \) based on the logarithm of the probability density functions, and a linear function \( u = ax+by > u_0 \) for optimal choices of \( a,b \). The quadratic function has a slightly lower expected cost.
What the @%&^ does this have to do with embedded systems?
We’ve been talking a lot about the mathematics of binary classification based on a two-test set of measurements with measurement \( x \) from test \( T_1 \) and measurement \( y \) from test \( T_2 \), where the probability distributions over \( (x,y) \) are a bivariate normal (Gaussian) distribution with slightly different values of mean \( (\mu_x, \mu_y) \) and covariance matrix \( \begin{bmatrix}\sigma_x{}^2 & \rho\sigma_x\sigma_y \cr \rho\sigma_x\sigma_y & \sigma_y{}^2\end{bmatrix} \) for two cases \( A \) and \( B \).
As an embedded system designer, why do you need to worry about this?
Well, you probably don’t need the math directly. It is worth knowing about different ways to use the results of the tests.
The more important aspect is knowing that a binary outcome based on continuous measurements throws away information. If you measure a sensor voltage and sound an alarm if the sensor voltage \( V > V_0 \) but don’t sound the alarm if \( V \le V_0 \) then this doesn’t distinguish the case of \( V \) being close to the threshold \( V_0 \) from the case where \( V \) is far away from the threshold. We saw this with idiot lights. Consider letting users see the original value \( V \) as well, not just the result of comparing it with the threshold. Or have two thresholds, one indicating a warning and the second indicating an alarm. This gives the user more information.
Be aware of the tradeoffs in choosing a threshold — and that someone needs to choose a threshold. If you lower the false positive rate, the false negative rate will increase, and vice versa.
Finally, although we’ve emphasized the importance of minimizing the chance of a false negative (in our fictional embolary pulmonism example, missing an EP-positive diagnosis may cause the patient to die), there are other costs to false positives besides inflicting unnecessary costs on users/patients. One that occurs in the medical industry is “alarm fatigue”, which is basically a loss of confidence in equipment/tests that cause frequent false positives. It is the medical device equivalent to the Aesop’s fable about the boy who cried wolf. One 2011 article stated it this.
One solution may be to abandon the ever-present beep and use sounds that have been designed to reduce alarm fatigue.
The real takeaway here is that the human-factors aspect of a design (how an alarm is presented) is often as important or even more important than the math and engineering behind the design (how an alarm is detected). This is especially important in safety-critical situations such as medicine, aviation, nuclear engineering, or industrial machinery.
References
- The accuracy of yes/no classification, University of Pennsylvania
- Sensitivity and Bias - an introduction to Signal Detection Theory, University of Birmingham
- Signal Detection Theory, New York University
- D-prime (signal detection) analysis, University of California at Los Angeles
- Calculation of signal detection theory measures, Stanislaw and Todorov, Behavior Research Methods, Instruments, & Computers, March 1999
Wrapup
We’ve talked in great detail today about binary outcome tests, where one or more continuous measurements are used to detect the presence of some condition \( A \), whether it’s a disease, or an equipment failure, or the presence of an intruder.
- False positives are when the detector erroneously signals that the condition is present. (consequences: annoyance, lost time, unnecessary costs and use of resources)
- False negatives are when the detector erroneously does not signal that the condition is present. (consequences: undetected serious conditions that may get worse)
- Choice of a threshold can trade off false positive and false negative rates
- Understanding the base rate (probability that the condition occurs) is important to avoid the base rate fallacy
- If the probability distribution of the measurement is a normal (Gaussian) distribution, then an optimal threshold can be chosen to maximize expected value, by using the logarithm of the probability density function (PDF), given:
- the first- and second-order statistics (mean and standard deviation for single-variable distributions) of both populations with and without condition \( A \)
- knowledge of the base rate
- estimated costs of all four outcomes (true positive detection, false positive, false negative, true negative)
- A binary outcome test can be examined in terms of a confusion matrix that shows probabilities of all four outcomes
- We can find the optimal threshold by knowing that if a measurement occurs at the optimal threshold, both positive and negative diagnoses have the same expected value, based the conditional probability of occurrence given the measured value
- Bayes’ Rule can be used to compute the conditional probability of \( A \) given the measured value, from its “converse”, the conditional probability of the measured value with and without condition \( A \)
- Idiot lights, where the detected binary outcome is presented instead of the original measurement, hide information from the user, and should be used with caution
- The distinguishability or separation coefficient \( d’ \) can be used to characterize how far apart two probability distributions are, essentially measuring the difference of the means, divided by an effective standard deviation
- If two tests are possible, with an inexpensive test \( T_1 \) offering a mediocre value of \( d’ \), and a more expensive test \( T_2 \) offering a higher value of \( d’ \), then the two tests can be combined to help reduce overall expected costs. We explored one situation (the fictional “embolary pulmonism” or EP) where the probability distribution was a known bivariate normal distribution. Five methods were used, in order of increasing effectiveness:
- Test \( T_1 \) only, producing a measurement \( x \), detecting condition \( A \) if \( x>x_0 \) (highest expected cost)
- Test \( T_2 \) only, producing a measurement \( y \), detecting condition \( A \) if \( y>y_0 \)
- Test \( T_1 \) and \( T_2 \), combined by a linear combination \( u = ax+by \), detecting condition \( A \) if \( u>u_0 \), with \( a \) and \( b \) chosen to maximize \( d’ \)
- Test \( T_1 \) and \( T_2 \), combined by a quadratic function \( \lambda(x,y) = a(x-\mu_x)^2 + b(x-\mu_x)(y-\mu_y) + c(y-\mu_y)^2 \) based on the log of the ratio of PDFs, detecting condition \( A \) if \( \lambda(x,y) > L \) (in our example, this has a slightly lower cost than the linear combination)
- Test \( T_1 \) used to detect three cases based on thresholds \( x_L \) and \( x_H \) (lowest expected cost):
- Diagnose absense of condition \( A \) if \( x < x_L \), no further test necessary
- Diagnose presence of condition \( A \) if \( x > x_H \), no further test necessary
- If \( x_L \le x \le x_H \), perform test \( T_2 \) and diagnose based on the measurements from both tests (we explored using the quadratic function in this article)
- An excessive false positive rate can cause “alarm fatigue” in which true positives may be missed because of a tendency to ignore detected conditions
Whew! OK, that’s all for now.
Previous post by Jason Sachs:
Wye Delta Tee Pi: Observations on Three-Terminal Networks
Next post by Jason Sachs:
Jaywalking Around the Compiler
-. | https://www.embeddedrelated.com/showarticle/1300.php | CC-MAIN-2020-05 | en | refinedweb |
20 Nov 2006
Table of Contents the URIs of the resources being requested and determine how best to satisfy those requests.
The best way to make this function in an interoperable way is to define a standard format for mapping system identifiers and URIs. The OASIS Entity Resolution Technical Committee is defining an XML representation for just such a mapping. These “catalog files” can be used to map public and system identifiers and other URIs to local files (or just other URIs).
The Resolver classes that are described in this article greatly simplify the task of using Catalog files to perform entity resolution. Many users will want to simply use these classes directly “out of the box” with their applications (such as Xalan and Saxon), but developers may also be interested in the JavaDoc API Documentation. The full documentation, current source code, and discussion mailing list are available from the Apache XML Commons project.
See the release notes.
The most important change in this release is the availability of both source and binary forms under a generous license agreement.
Other than that, there have been a number of minor bug fixes and the introduction
of system properties in addition to the
CatalogManager.properties
file to control the resolver.
The problems associated with system identifiers (and URIs in general) arise in several ways:.1.2//EN" "">
As soon as I distribute this document, I immediately begin getting error reports from customers who can't read the document because they don't have DocBook installed at the location identified by the URI in my document.
Or I remember to change the URI before I publish the document:
<!DOCTYPE article PUBLIC "-//OASIS//DTD DocBook XML V4.1.2//EN" "">
And the next time I try to edit the document, I get errors because I happen to be working on my laptop on a plane somewhere and can't get to the net.
Just as often, I get tripped up this way: I'm working collaboratively with a colleague. She's created initial drafts of some documents that I'm supposed to review and edit. So I grab them and find that I can't open or publish them because I don't have the same network connections she has or I don't have my applications installed in the same place. And if I change the system identifiers so they work on my system, she has the same problems when I send them back to her.
These problems aren't limited to editing applications. If I write a special stylesheet for formatting our collaborative document, it will include some reference to the “main” stylesheet:
<xsl:import
But this won't work on my colleague's machine because she has the main stylesheet installed somewhere else.
Public identifiers offer an effective solution to this problem, at least for documents. They provide global, unique names for entities independent of their storage location. Unfortunately, public identifiers aren't used very often because many users find that they cannot rely on applications resolving them in an interoperable manner.
For XSLT, XML Schemas, and other applications that rely on URIs without providing a mechanism for associating public identifiers with them, the situation is a little more irksome, but it can still be addressed using a URI Resolver.
In some contexts, it's more useful to refer to a resource by name than by address. If I want the version 3.1 of the DocBook DTD, or the 1911 edition of Webster's dictionary, or The Declaration of Independence,].
Public identifiers are part of XML 1.0. They can occur in any form of external entity declaration. They allow you to give a globally unique name to any entity. For example, the XML version of DocBook V4.1.2 is identified with the following public identifier:
-//OASIS//DTD DocBook XML V4.1.2/.1.2.
URNs are a form of URI. Like public identifiers, they give a location-neutral, globally unique name to an entity. For example, OASIS might choose to identify the XML version of DocBook V4.1.2 with the following URN:
urn:oasis:names:specification:docbook:dtd:xml:4.1.2
Like a public identifier, a URN can now and forever refer to a specific entity in a location-independent manner.
Public identifiers don't fit very well into the web architecture
(they are not, for example, always valid URIs). This problem can be
addressed by the
publicid URN namespace defined by
RFC 3151.
This namespace allows public identifiers to be easily represented as URNs. The OASIS XML Catalog specification accords special status to URNs of this form so that catalog resolution occurs in the expected way..1.2, if you have a local copy.
There are a few possible resolution mechanisms:
The application just “knows”. Sure, it sounds a little silly, but this is currently the mechanism being used for namespaces. Applications know what the semantics of namespaced elements are because they recognize the namespace URI.
OASIS Catalog files provide a mechanism for mapping public and system identifiers, allowing resolution to both local and distributed resources. This is the resolution scheme we're going to consider for the balance of this column.
Many other mechanisms are possible. There are already a few for URNs, including at least one built on top of DNS, but they aren't widely deployed.
Catalog files are straightforward text files that describe a mapping from names to addresses. Here's a simple one:
Example 1. An Example Catalog File
<catalog xmlns="urn:oasis:names:tc:entity:xmlns:xml:catalog"> <public publicId="-//OASIS//DTD XML DocBook V4.1.2//EN" uri="docbook/xml/docbookx.dtd"/> <system systemId="urn:x-oasis:docbook-xml-v4.1.2" uri="docbook/xml/docbookx.dtd"/> <delegatePublic publicIdStartString="-//Example//" catalog=""/> </catalog> “”
for any public identifier that begins with “-//Example//”.
The advantage of delegate in this case is that I don't have to parse that
catalog file unless I encounter a public identifier that I reasonably expect
to find there.
The OASIS Entity Resolution Technical Committee is actively defining the next generation XML-based catalog file format. When this work is finished, it is expected to become the official XML Catalog format. In the meantime, the existing OASIS Technical Resolution TR9401 format is the standard.
OASIS XML Catalogs are being defined by the Entity Resolution Technical Committee. This article describes the 01 Aug 2001 draft. Note that this draft is labelled to reflect that it is “not an official committee work product and may not reflect the consensus opinion of the committee.”
The document element for OASIS XML Catalogs is
catalog. The official namespace name for OASIS XML
Catalogs is
“
urn:oasis:names:tc:entity:xmlns:xml:catalog”.
There are eight elements that can occur in an XML Catalog:
group,
public,
system,
uri,
delegatePublic,
delegateSystem,
delegateURI, and
nextCatalog:
<catalog
prefer="public|system"
xml:
The
catalog element is the root of
an XML Catalog.
The
prefer setting
determines whether or not public identifiers specified in the catalog
are to be used in favor of system identifiers supplied in the
document. Suppose you have an entity in your document for which both a
public identifier and a system identifier has been specified, and the
catalog only contains a mapping for the public identifier (e.g., a
matching
public catalog entry). If the current
value of
prefer is
“public”, the URI supplied in the matching
public catalog entry will be used. If it is
“system”, the system identifier in the document will be
used. (If the catalog contained a matching
system
catalog entry giving a mapping for the system identifier, that mapping
would have been used, the public identifier would never have been
considered, and the setting of override would have been
irrelevant.)
Generally, the purpose of catalogs is to
override the system identifiers in XML documents, so
prefer should
usually be “public” in your catalogs.
The
xml:base URI is used to
resolve relative URIs in the catalog as described in the
XML Base specification.
<group
prefer="public|system"
xml:
The
group element serves merely as
a wrapper around one or more other entries for the purpose of
establishing the preference and base URI settings for those
entries.
<public publicId="
pubid" uri="
systemuri"/>
Maps the public identifier
pubid to the
system identifier
systemuri.
<system systemId="
sysid" uri="
systemuri"/>
Maps the system identifier
sysid to the
alternate system identifier
systemuri.
<uri name="
uri" uri="
alternateuri"/>
The
uri entry maps a
uri to an
alternateuri. This mapping, as might be performed
by a JAXP URIResolver, for example, is independent of system and public
identifier resolution.
<delegatePublic publicIdStartString=",
pubid-prefix" catalog="
cataloguri"/>
<delegateSystem systemIdStartString=",
sysid-prefix" catalog="
cataloguri"/>
<delegateURI uriStartString="
uri-prefix" catalog="
cataloguri"/>
The delegate entries specify that identifiers beginning with the
matching prefix should be resolved using the catalog specified by the
cataloguri. If multiple delegate entries
of the same kind match, they will each be searched, starting with the
longest prefix and continuing with the next longest to the
shortest.
The delegate entries differs from the
nextCatalog entry in the following way: alternate
catalogs referenced with a
nextCatalog entry are parsed
and included in the current catalog. Delegated catalogs are only
considered, and consequently only loaded and parsed, if
necessary. Delegated catalogs are also used instead
of the current catalog, not as part of the current
catalog.
<rewriteSystem systemIdStartString=",
sysid-prefix" rewritePrefix="
new-prefix"/>
<rewriteURI uriStartString="
uri-prefix" rewritePrefix="
new-prefix"/>
Supports generalized rewriting of system identifiers and URIs. This allows all of the URI references to a particular document (which might include many different fragment identifiers) to be remapped to a different resource).
<nextCatalog catalog="
cataloguri"/>
Adds the catalog file specified by the
cataloguri
to the end of the current catalog. This allows one catalog to refer to another.
These catalogs are officially defined by OASIS Technical Resolution TR9401.
A Catalog is a text file that contains a sequence of entries. Of the 13 types of entries that are possible, only six are commonly applicable in XML systems: BASE, CATALOG, OVERRIDE, DELEGATE, PUBLIC, and SYSTEM:
uri
Catalog entries can contain relative URIs. The BASE entry changes the base URI for subsequent relative URIs. The initial base URI is the URI of the catalog file.
In XML Catalogs, this
functionality is provided by the closest applicable
xml:base attribute, usually on the
surrounding
catalog
or
group
element.
cataloguri
This entry serves the same purpose as the
nextCatalog entry
in XML Catalogs.
YES|NO
This entry enables or disables overriding of system identifiers for subsequent entries in the catalog file.
In XML Catalogs, this
functionality is provided by the closest applicable
prefer attribute on the
surrounding
catalog
or
group
element.
An override value of “yes” is equivalent to “prefer="public"”.
pubid-prefix
cataloguri
This entry serves the same purpose as the
delegate entry
in XML Catalogs.
pubid
systemuri
This entry serves the same purpose as the
public entry
in XML Catalogs.
sysid
systemuri
This entry serves the same purpose as the
system entry
in XML Catalogs.
Resolution is performed in roughly the following way:
If a system entry matches the specified system identifier, it is used.
If no system entry matches the specified system identifier, but a rewrite entry matches, it is used.
If a public entry matches the specified public identifier
and either
prefer
is public or no system identifier is provided,
it is used.
If no exact match was found, but it matches one or more of the partial identifiers specified in delegate entries, the delegated catalogs are searched for a matching identifier.
For a more detailed description of resolution semantics, including the treatment of multiple catalog files and the complete rules for delegation, consult the XML Catalog standard.
The Resolver classes uses either Java system properties or a
standard Java properties file to establish an initial environment. The
property file, if it is used, must be called
CatalogManager.properties and must be
somewhere on your
CLASSPATH. The following properties
are supported:
xml.catalog.files; CatalogManager.
xml.catalog.prefer; CatalogManager property
prefer
The initial prefer setting, either
public
or
system.
xml.catalog.verbosity; CatalogManager property
verbosity
An indication of how much status/debugging information you want to receive. The value is a number; the larger the number, the more information you will receive. A setting of 0 turns off all status information.
xml.catalog.staticCatalog; CatalogManager property
static-catalog
In the course of processing, an application may parse
several XML documents. If you are using the built-in
CatalogResolver, this option controls whether or
not a new instance of the resolver is constructed for each parse.
For performance reasons, using a value of
yes, indicating
that a static catalog should be used for all parsing, is probably best.
xml.catalog.allowPI; CatalogManager property
allow-oasis-xml-catalog-pi
This setting allows you to toggle whether or not the
resolver classes obey the
<?oasis-xml-catalog?>
processing instruction.
xml.catalog.className; CatalogManager property
catalog-class-name
If you're using the convenience classes
org.apache.xml.resolver.tools.*), this setting
allows you to specify an alternate class name to use for the underlying
catalog.
relative-catalogs
If
relative-catalogs is
yes,
relative catalogs in the
catalogs property will be left relative;
otherwise they will be made absolute
with respect to the base URI of the
CatalogManager.properties
file. This setting has no effect on catalogs loaded from the
xml.catalogs.files system property (which are always returned
unchanged).
xml.catalog.ignoreMissing
By default, the resolver will issue warning messages if it
cannot find a
CatalogManager.properties file, or if resources
are missing in that file. However if either
xml.catalog.ignoreMissing is
yes, or
catalog files are specified with the
xml.catalog.catalogs system property, this warning will
be suppressed.
My
CatalogManager.properties file looks like
this:
Example 2. Example CatalogManager.properties File
#CatalogManager.properties verbosity=1 relative-catalogs=yes # Always use semicolons in this list catalogs=./xcatalog;/share/doctypes/catalog;/share/doctypes/xcatalog prefer=public static-catalog=yes allow-oasis-xml-catalog-pi=yes catalog-class-name=org.apache.xml.resolver.Resolver
A number of popular applications provide easy access to catalog resolution:
Recent development versions of Xalan include new command-line
switches for setting the resolvers. You can use them directly with
the
org.apache.xml.resolver.tools classes:
-URIRESOLVER org.apache.xml.resolver.tools.CatalogResolver -ENTITYRESOLVER org.apache.xml.resolver.tools.CatalogResolver
Similarly, Saxon supports command-line access to the resolvers:
-x org.apache.xml.resolver.tools.ResolvingXMLReader -y org.apache.xml.resolver.tools.ResolvingXMLReader -r org.apache.xml.resolver.tools.CatalogResolver
The
-x class is used to read source documents,
the
-y class is used to read stylesheets.
To use XP, simply use the included
org.apache.xml.xp.xml.sax.Driver class instead of
the default XP driver.
Similarly, for XT, use the
org.apache.xml.xt.xsl.sax.Driver class.
If you work with Java applications using a parser that supports
the SAX1
Parser interface or the SAX2
XMLReader interface, adding Catalog support to your
applications is a snap. The SAX interfaces
include an
entityResolver hook designed to provide
an application with an opportunity to do this sort of indirection. The
Resolver classes implements the full
OASIS Catalog semantics and provide an appropriate class that
implements the SAX
entityResolver interface.
All you have to do is setup a
org.apache.xml.resolver.tools.CatalogResolver
on your parser's
entityResolver hook. The code listing
in Example 3, “Adding a CatalogResolver to Your Parser” demonstrates how straightforward this is:
Example 3. Adding a CatalogResolver to Your Parser
import org.apache.xml.resolver.tools.CatalogResolver; ... CatalogResolver cr = new CatalogResolver(); ... yourParser.setEntityResolver(cr)
The system catalogs are loaded from the
CatalogManager.properties file on your
CLASSPATH.
(For all the
gory details about these classes, consult the
API documentation.) You can explicitly parse your own catalogs (perhaps
taken from command line arguments or a Preferences dialog) instead of or in
addition to the system catalogs.
The Resolver distribution includes a couple of test programs, resolver and xparse, that you can use to see how this all works.
The resolver application simply performs a catalog lookup and returns the result. Given the following catalog:
Example 4. An Example XML Catalog File
<catalog xmlns="urn:oasis:names:tc:entity:xmlns:xml:catalog"> <public publicId="-//Example//DTD Example V1.0//EN" uri="example.dtd"/> </catalog>
A demonstration of public identifier resolution can be achieved like this:
Example 5. Resolving Identifiers
$ java org.apache.xml.resolver.apps.resolver -d 2 -c example/catalog.xml \ -p "-//Example//DTD Example V1.0//EN" public Loading catalog: ./catalog Loading catalog: /share/doctypes/catalog Resolve PUBLIC (publicid, systemid): public id: -//Example//DTD Example V1.0//EN Loading catalog: file:/share/doctypes/entities.cat Loading catalog: /share/doctypes/xcatalog Loading catalog: example/catalog.xml Result: file:/share/documents/articles/sun/2001/01-resolver/example/example.dtd
The xparse command simply sets up a catalog resolver and then parses a document. Any external entities encountered during the parse are resolved appropriately using the catalogs provided.
In order to use the program, you must have the
resolver.jar file on your
CLASSPATH and you must be using JAXP. In the examples that
follow, I've already got these files on my
CLASSPATH.
The file we'll be parsing is shown in Example 6, “An xparse Example File”.
Example 6. An xparse Example File
<!DOCTYPE example PUBLIC "-//Example//DTD Example V1.0//EN" ""> <example> <p>This is just a trivial example.</p> </example>
First let's look at what happens if you try to parse this
document without any catalogs. For this example, I deleted the
catalogs entry on my
CatalogManager.properties file. As expected,
the parse fails:
Example 7. Parsing Without a Catalog
$ java org.apache.xml.resolver.apps.xparse -d 2 example.xml Attempting validating, namespace-aware parse Fatal error:example.xml:2:External entity not found: "". Parse failed with 1 error and no warnings.
With an appropriate catalog file, we can map the public identifier to a local copy of the DTD. We could have mapped the system identifier instead (or as well), but the public identifier is probably more stable.
Using a command-line option to specify the catalog, I can now successfully parse the document:
Example 8. Parsing With a Catalog
$ java org.apache.xml.resolver.apps.xparse -d 2 -c catalog.xml example.xml Loading catalog: catalog.xml Attempting validating, namespace-aware parse Resolved public: -//Example//DTD Example V1.0//EN file:/share/documents/articles/sun/2001/01-resolver/example/example.dtd Parse succeeded (0.32) with no errors and no warnings.
The additional messages in each of these examples arise as a
consequence of the debugging option,
-d 2.
In practice, you can make resolution silent.
We hope that these classes become a standard part of your toolkit. Incorporating this code allows you to utilize public identifiers in XML documents with the confidence that you will be able to move those documents from one system to another and around the Web.
[1] It is technically possible to use a proxy to transparently cache remote resources, thus making the cached resources available even when the real hosts are unreachable. In practice, this requires more technical skill (and system administration access) than many users have available. And I don't know of any such proxies that can be configured to provide preferential caching to the specific resources that are needed. Without such preferential treatment, its difficult to be sure that the resources you need are actually in the cache. | https://xerces.apache.org/xml-commons/components/resolver/resolver-article.html | CC-MAIN-2020-05 | en | refinedweb |
Starting with v6.0, the Agent can ingest metrics with a Unix Domain Socket (UDS) as an alternative to UDP when running on Linux systems.
While UDP works great on
localhost, it can be a challenge to set up in containerized environments. Unix Domain Sockets allow you to establish the connection with.
To set up DogStatsD, enable the DogStatsD server through the
dogstatsd_socket parameter. Then, configure the DogStatsD client in your code.
To enable the Agent DogStatsD server:
Edit the Agent’s main configuration file to set
dogstatsd_socket to the path where DogStatsD should create its listening socket:
## @param dogstatsd_socket - string - optional - default: "" ## Listen for Dogstatsd metrics on a Unix Socket (*nix only). Set to a valid filesystem path to enable. # dogstatsd_socket: "/var/run/datadog/dsd.socket"
Note: You can also set the socket path with the
DD_DOGSTATSD_SOCKET environment variable for the container Agent.
The following official DogStatsD client libraries natively support UDS traffic. Refer to the library’s documentation on how to enable UDS traffic. Note: As with UDP, enabling client-side buffering is highly recommended to improve performance on heavy traffic:
For guidelines on creating additional implementation options, refer to the datadog-agent github wiki. are tagged by the same container tags as Autodiscovery metrics.
Note:
container_id,
container_name, and
pod_name tags are not added to avoid creating too many custom metrics.
To use origin detection:
Enable the
dogstatsd_origin_detection option in your Agent’s main configuration file:
## @param dogstatsd_origin_detection - boolean - optional - default: false ## When using Unix Socket, DogStatsD can tag metrics with container metadata. ## If running DogStatsD in a container, host PID mode (e.g. with --pid=host) is required. # dogstatsd_origin_detection: true
Note: For the containerized Agent, set the environment variable
DD_DOGSTATSD_ORIGIN_DETECTION to true.
When running inside a container, DogStatsd needs to run in the host’s PID namespace for origin detection to work reliably. Enable this with the Docker
--pid=host flag.
Note: This is supported by ECS with the parameter
"pidMode": "host" in the task definition of the container. This option is not supported in Fargate. For more information, see the AWS documentation. | https://docs.datadoghq.com/ja/developers/dogstatsd/unix_socket/ | CC-MAIN-2020-05 | en | refinedweb |
acl_get_file()
Get the ACL for a given path
Synopsis:
#include <sys/acl.h> acl_t acl_get_file( const char *path_p, acl_type_t type );
Since:
BlackBerry 10.0.0
Arguments:
- path_p
- The path that you want to get the ACL for.
- type
- The type of ACL; this must currently be ACL_TYPE_ACCESS.
Library:
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
Description:
The acl_get_file() function gets the ACL associated with the given file or directory, copies it into working storage, and returns a pointer to that storage. When you're finished with the ACL, you should call acl_free() to release it.
The ACL in working storage is independent of the file or directory's ACL. Changes that you make to the copy in working storage don't affect the file or directory's ACL.
Errors:
- EACCES
- Search permission was denied for a component of the path prefix, or the object exists and the process doesn't have the appropriate access rights.
- EINVAL
- The type argument isn't ACL_TYPE_ACCESS.
- ENAMETOOLONG
- The length of the path_p argument exceeds PATH_MAX.
- ENOENT
- The named object doesn't exist, or path_p is an empty string.
- ENOMEM
- There wasn't enough memory available to create the ACL in working storage.
- ENOTDIR
- A component of the path prefix isn't a directory.
Classification:
This function is based on the withdrawn POSIX draft P1003.1e.
Last modified: 2014-11-17
Got questions about leaving a comment? Get answers from our Disqus FAQ.comments powered by Disqus | https://developer.blackberry.com/native/reference/core/com.qnx.doc.neutrino.lib_ref/topic/a/acl_get_file.html | CC-MAIN-2020-05 | en | refinedweb |
- Part 1: Introduction
- Part 2: Effects
Let's imagine you want to write a function which writes to a file in Node. Here's a simple example:
import * as fs from "fs"; function writeHelloWorld(path: string) { fs.writeFileSync(path, "Hello world!"); } writeHelloWorld("my-file.md");
What's the type of our function
writeHelloWorld?
fs.writeFileSync returns
void, so naturally so does
writeHelloWorld.
But let's think about this for a second. What if we took the same function and made it empty? Like this:
function writeHelloWorld(path: string) { // Nothing to see here! }
It's got the exact same type
(path: string) => void. But of course they do very different things. What the first example does - writing a file - is called an effect. It's certainly not a pure function.
If you're familiar with React hooks, this is exactly the same concept as
useEffect, which you use for things like fetching data (which is very much an effect).
We're going to run on the assumption here that writing pure functions is good, and having side effects is bad.
So how could we make
writeHelloWorld pure? Is this possible?
How about we make it lazy by having
writeHelloWorld return a function?
import * as fs from "fs"; function writeHelloWorld(path: string) { return () => fs.writeFileSync(path, "Hello world!"); } const runEffects = writeHelloWorld("my-file.md");
Is this code now pure?
- no side effects ✅
- same input always produces same output ✅
Seems pure to me! But it doesn't do anything. We also need to add:
runEffects();
Only this part of our code is impure, because without that, nothing would happen.
Is this cheating? Sort of, but it works.
In fact we can make our entire program pure, despite it being full of effects like file writes and random numbers all over the place. So long as every function with effects is lazy, we can combine all of our effects together into one lazy function - like a big red button to run the program:
// WARNING: do not call this function unless you're sure what you're doing! function runMyPureProgram() { return () => { writeSomeFiles(); getSomeRandomNumbers(); getTodaysDate(); } }
Calling
runMyPureProgram() is impure, yet the program itself is pure.
But how do we ensure our functions are pure? If I look at
getSomeRandomNumbers and it returns
() => number[], how do I know if it has an effect or not? If only there was some way the compiler could tell us this... 🤔
Well it's called Type Script for a reason! Let's imagine a new type:
type Effect<T> = () => T;
What we say is, any impure function with effects must return the type
Effect.
T is the return value - for example, if our effect was to read a file as a string it could be
Effect<string>.
We can write:
function writeHelloWorld(path: string): Effect<void> { return () => fs.writeFileSync(path, "Hello world!"); }
Now, just by looking at the type signature of
writeHelloWorld, we know it's going to be doing some effectful things. In this case,
T is
void since it doesn't return a value.
So if I start adding new impure things in my code, TypeScript will complain until I actually handle the effects correctly at the end of the program.
Through some lazy functions and a new type called
Effect, we've both kept our program pure and we know where the impure effectful things are hiding. 🤯
If you're like me, at this point you might be thinking: "This is great! Wait - what's the point?". It seems like we still have a bunch of impure stuff littered around our codebase, and we've now made the code more awkward than it was before. All in the name of "purity".
Let's stop for a second for a quiz. Imagine this contrived example:
function funcA(): number { return funcB(); } function funcB(): number { return funcC(); } function funcC(): number { return 5; }
We want to change
funcC to return a random number. I'll give you three ways we could do this, which one is the 'best'?
1. The impure way:
function funcC(): number { return Math.random(); }
Easy! Nothing else to change, let's go home now.
2. Dependency injection:
function funcA(randomGenerator: () => number): number { return funcB(randomGenerator); } function funcB(randomGenerator: () => number): number { return funcC(randomGenerator); } function funcC(randomGenerator: () => number): number { return randomGenerator(); }
Our functions are still nice and pure, but we had to add an argument to them all, which is a lot of typing...
3. Effects
function funcA(): Effect<number> { return funcB(); } function funcB(): Effect<number> { return funcC(); } function funcC(): Effect<number> { return () => Math.random(); }
Everything's still pure but we had to change the return type of everything to make it lazy.
So I'll ask again, which one's the 'best'?
I would say typing effects are a blessing and a curse. It's exactly because they require you to change your function signatures you're forced to program in a way where effects are kept to a minimum in your code. And you're more likely to put them all in one place.
It's like immutability - it can be more annoying to do, but can provide a cleaner codebase overall. Whether the short-term pain is worth the long-term gain is always difficult to say though.
Even if you're now convinced to use the
Effect type, this still isn't perfect. There's nothing in TypeScript which forces us to add an
Effect type for effects, we just have to be vigilant about it. If you want to be forced to do it, that's when you head to pure a functional language like Haskell.
Introducing: IO
The fp-ts library already has a type just like
Effect, and its name is
IO. In fact, its type is exactly as we defined
Effect.
// taken from source code export interface IO<A> { (): A }
The benefit of using
IO is it comes with a lot of helper functions for handling effects. Plus, it provides functions for effectful things (like generating random numbers) which are already typed as
IO. Here are some simple examples:
Add a number
Let's start here:
import { IO } from "fp-ts/lib/IO"; import { random } from "fp-ts/lib/Random"; function funcA(): IO<number> { return funcB(); } function funcB(): IO<number> { // importing random from fp-ts is the same as // return () => Math.random(); return random; }
Imagine we want to add 1 in
funcA:
function funcA(): IO<number> { return funcB() + 1; }
Oops! We get the error:
Operator '+' cannot be applied to types 'IO<number>' and 'number'.
We could change
funcA to this:
function funcA(): IO<number> { return () => funcB()() + 1; }
But that's annoying and plain ugly. Instead fp-ts gives us
io.map:
import { IO, io } from "fp-ts/lib/IO"; function funcA(): IO<number> { return io.map(funcB(), (x) => x + 1); }
Much like you call
.map on an array, you can
map something of type
IO too.
Logging a random number
We can use
io.chain to combine two effects into a single effect:
import { IO, io } from "fp-ts/lib/IO"; import { random } from "fp-ts/lib/Random"; import { log } from "fp-ts/lib/Console"; function logRandomNumber(): IO<void> { return io.chain(random, log); } const program = logRandomNumber(); // this logs a random number when called program();
In this way we can combine all of our effects into one
program, which we then call right at the very end.
There's much more to learn that we haven't covered (I haven't even mentioned error handling), so do check out these resources if you want to find out more:
Discussion (2)
Surely at times this would be unnecessary closure allocation?
Interesting point! I guess you need to weigh up developer productivity vs performance. In the vast majority of cases I doubt this would add any noticeable difference to performance. | https://dev.to/edbentley/mind-bending-functional-programming-with-typescript-part-2-40c9 | CC-MAIN-2021-10 | en | refinedweb |
Very nice, Can you turn the money feature off ? for milsim and such.
Great Stuff!
Fantastic! This build has been the most fun to play with. thanks for the upload!
is there a way to decrease the amount of time to build like instead of having to hit the build option 10 times could i set it to like 5..... or 2 build and how do i add other items to Build/buy
Total comments : 4, displayed on page: 4
null = [] execVM "Loli_Defense\LD_Init.sqf";
#include "Loli_Defense\LD_GUI_Defines.hpp"
#include "Loli_Defense\LD_GUI_Dialogs.hpp"
this setVariable ["materials", amount_of_material, true];
this setVariable ["cash", amount_of_cash,! | https://www.armaholic.com/page.php?id=24150 | CC-MAIN-2021-10 | en | refinedweb |
Touch Support
The Telerik UI for WinForms suite provides full multi-touch support. All controls in the suite are exposing several events which grant the developer the ability to easily handle gestures on a touch devices. This functionality is currently supported under Windows7 or newer. Some of the controls in our suite have built-in functionality that responds to touch gestures. For example, you can use the Pan gesture to scroll through the RadGridView’s rows, group by a column or change the order of its columns. Similar functionality is available out-of-the-box for RadTreeView, RadListView, RadPropertyGrid, RadListControl, RadCarousel and RadCommandBar. Additionally, the developer can use the gesture events to implement his custom logic.
Touch Events in RadControls
To enable or disable a gesture,use the EnableGesture and DisableGesture functions passing a member of the GestureType enumerator. This method should be executed in the constructor of a new control:
All
Pan
Rotate
Zoom
TwoFingerTap
PressAndTap
Cascade layout example
public class MyButton : RadButton { public MyButton() { this.EnableGesture(GestureType.All); this.DisableGesture(GestureType.Zoom); } }
Public Class MyButton Inherits RadButton Public Sub New() Me.EnableGesture(GestureType.All) Me.DisableGesture(GestureType.Zoom) End Sub End Class
An explanation of the different gestures can be found in this MSDN article.
You can use the following events to handle gesture events:
PanGesture: Fires when the user slides with his finger across the area of the control.
ZoomGesture: Fires when the user slides with his two fingers in opposite directions.
RotateGesture: Fires when the user slides with his two fingers in a circular direction.
TwoFingerTapGesture: Fires when the user taps the screen with his two fingers at the same time.
PressAndTapGesture: Fires when the user has pressed the screen with a finger and taps with a second finger.
All these events provide event arguments that inherit from the GestureEventArgs type, hence the share the following properties:
IsBegin: Indicates that the gesture is starting.
IsEnd: Indicates that the gesture is ending.
IsInertia: Indicates that the event is caused by inertia.
Location: Indicates the location in control coordinates at which the gesture has occurred.
Handled: Indicates if the event has already been handled by some of the elements in the control.
The inheritors of this type also provide gesture-specific arguments like Offset, ZoomFactor, Angle etc.
Touch Events in RadItems
All the above mentioned events are also valid for all RadItems. This means you can use them for different items in RadRibbonBar, RadCommandBar, RadMenu, etc.
Example of Using Touch Events
The following example will demonstrate how we can use this functionality to drag, rotate and resize RadButtonElement within a simple panel:
public class CustomPanel : RadControl { public class CustomPanelLayout : Telerik.WinControls.Layouts.LayoutPanel { } public CustomPanel() { this.EnableGesture(GestureType.All); } CustomPanelLayout m_layout; RadButtonElement button; protected override void CreateChildItems(RadElement parent) { base.CreateChildItems(parent); m_layout = new CustomPanelLayout(); parent.Children.Add(m_layout); button = new RadButtonElement(); button.AutoSize = false; button.Size = new Size(100, 100); button.Location = new Point(100, 100); button.Text = "RadButtonElement"; this.m_layout.Children.Add(button); button.PanGesture += new PanGestureEventHandler(button_PanGesture); button.ZoomGesture += new ZoomGestureEventHandler(button_ZoomGesture); button.RotateGesture += new RotateGestureEventHandler(button_RotateGesture); } void button_RotateGesture(object sender, RotateGestureEventArgs e) { button.AngleTransform -= (float)(e.Angle * 180D / Math.PI); } void button_ZoomGesture(object sender, ZoomGestureEventArgs e) { button.ScaleTransform = new SizeF( (float)(button.ScaleTransform.Width * e.ZoomFactor), (float)(button.ScaleTransform.Height * e.ZoomFactor)); } void button_PanGesture(object sender, PanGestureEventArgs e) { button.Location = new Point(button.Location.X + e.Offset.Width, button.Location.Y + e.Offset.Height); } }
Public Class CustomPanel Inherits RadControl Public Class CustomPanelLayout Inherits Telerik.WinControls.Layouts.LayoutPanel End Class Public Sub New() Me.EnableGesture(GestureType.All) End Sub Private m_layout As CustomPanelLayout Private button As RadButtonElement Protected Overrides Sub CreateChildItems(ByVal parent As RadElement) MyBase.CreateChildItems(parent) m_layout = New CustomPanelLayout() parent.Children.Add(m_layout) button = New RadButtonElement() button.AutoSize = False button.Size = New Size(100, 100) button.Location = New Point(100, 100) button.Text = "RadButtonElement" Me.m_layout.Children.Add(button) AddHandler button.PanGesture, AddressOf button_PanGesture AddHandler button.ZoomGesture, AddressOf button_ZoomGesture AddHandler button.RotateGesture, AddressOf button_RotateGesture End Sub Private Sub button_RotateGesture(ByVal sender As Object, ByVal e As RotateGestureEventArgs) button.AngleTransform -= CSng(e.Angle * 180.0R / Math.PI) End Sub Private Sub button_ZoomGesture(ByVal sender As Object, ByVal e As ZoomGestureEventArgs) button.ScaleTransform = New SizeF(CSng(button.ScaleTransform.Width * e.ZoomFactor), CSng(button.ScaleTransform.Height * e.ZoomFactor)) End Sub Private Sub button_PanGesture(ByVal sender As Object, ByVal e As PanGestureEventArgs) button.Location = New Point(button.Location.X + e.Offset.Width, button.Location.Y + e.Offset.Height) End Sub End Class
Thanks to the code above, the end-user will be able to do the following operations with his/her fingers:
Similar functionality is also used in the PhotoAlbum demo application | https://docs.telerik.com/devtools/winforms/telerik-presentation-framework/touch-support | CC-MAIN-2021-10 | en | refinedweb |
More factors, more variance…explained
Want to share your content on python-bloggers? click here.. Results were interesting, but begged the question why we were applying predominantly equity factors to portfolios of diverse assets: bonds, commodities, and real estate in addition to stocks. The simple answer: it was easy to get the data and the factors are already relatively well known.
In our second investigation, we built a risk factor model using economic variables that explained surprisingly little of the four portfolios’ variance. We admit our choice of macro factors wasn’t terribly scientific. But we felt, intuitively, they captured a fair amount of the US economy.1 Unfortunately, the macro factor model did an even worse job explaining portfolio variance than the F-F model.
We hypothesized that some of the problems with the macro factor model—besides data issues around timing and frequency—were that not all of the series we used were leading indicators. In this post, we’ll re-run the model, but with putative leading factors and then see what we can do to improve the model, if at all. Let’s begin!
We start by pulling the data. In the previous post, we used PMI, the unemployment rate, personal consumption expenditures, housing permits, consumer sentiment, and yield spreads (e.g. ten-year less two-year Treasuries, and Baa-rated less Aaa-rated corporates). In this post, we’ll use PMI, initial unemployment claims, the leading index2, housing permits, monetary base, and yield spreads. If these substitutions don’t match your view of true leading indicators, let us know at the email address below. We’re not economists, nor do we pretend to be experts of the dismal science. The variables we use are based on our search into generally accepted leading indicators.
First, here’s the spaghetti graph of the variables transformed to month-over-month sequential changes. We exclude yield spreads since they are not reported as a sequential change, and, if we were to transform them, they would overwhelm the other data.
Now, we’ll use that data to run our \(R^{2}\)s analysis on the various asset classes.
Pretty disappointing. The high \(R^{2}\) for real estate is odd, but we’re not going to dwell on it. We believe working through a full, normalized data series will prove more fruitful. Nonetheless, we’ll show how this macro factor model explains the variance of the four portfolios. We’re skipping over showing how we calculate factor exposures and idiosyncratic risk; you can find that in the code below and the first post. Here’s the explained variance graph.
This is actually worse than the previous data in terms of explanatory power. Perhaps our intuition wasn’t so bad after all! Let’s move on.
Next we’ll normalize the macro variables and regress them against forward returns. Not knowing which combination of look back and look forward windows is the best, we perform a grid search of rolling three-to-twelve month look back normalizations and one-to-twelve month forward returns, equivalent to 120 different combinations. We sort the results by the highest \(R^{2}\) for stocks and bonds and show the top five below in a table.
We’re not sure why three-to-five month look backs on eight month forward returns produce the highest explanatory power. But lets run with the five-month look back and eight month forward (5-8 model). Next we’ll show the factor \(\beta\)s by asset class.
Some of the size effects are reasonable in terms of magnitude and direction: monetary base increasing should be positive for stock returns while corporate yield spreads widening should be negative. But others aren’t so much. An increasing monetary base should suggest rising inflation which should be positive for gold. Yet, it appears to have a negative impact for the yellow metal. However, widening corporate spreads, presaging rising risk, should see some correlation with that gilded commodity.
Let’s see how much the 5-8 model explains the variance of the four portfolios.
Certainly better than the base model and better than the 9-9 model from the previous post. But can we do better?
We’re going to make a broad brush generalization that we won’t try to prove, but we think is generally right. If you think we’re wrong, by all means reach out at the email address below. Here it is. For most investors, most of the time, stocks will be the bulk of the portfolio. They may not be the majority, but they’re likely to be at least a plurality for say 70% of liquid investment portfolios.3 More importantly, unless you’re invested in an alternative asset class or using an arcane investment strategy, stocks will likely produce the bulk of the returns and volatility. Given this gross generalization, perhaps we should include factors that explain stock returns better than the macro variables. Bring back Fama-French!
We add the Fama-French factors to the macro variables and run a grid search as above. Note, we’re excluding the market risk premium as well as risk free rate. Hence, as in the foregoing analysis, we’re analyzing raw returns, rather than risk premia. The grid search table is below.
The look back period clusters around nine to eleven, with returns five-months in the future dominating the top five spots. We’ll go with the ten-month look back and five-month forward (10-5 model).
Explained variance increased more than five-to-seven percentage points on average with the addition of the F-F factors.
For a final twist, let’s adjust the assets to a risk premium type return (return less the risk free rate) and include the market risk premium. We have to be careful, however. The F-F market risk premium is highly correlated with our stock risk premium.4 So if we included the market risk premium we’d explain almost all of the variance in our stock returns on a coincident basis. On a lagged basis, we’d be incorporating a autoregressive model for stock returns that wouldn’t be present for the other asset classes.
Should we exclude the market risk premium? We think not. Conceptually, the market risk premium is supposed to capture most of the systematic risk not identified by the other factors. This means that even if it is primarily equity market-based, it would likely explain some of the variance of other asset classes, given how much the stock market reflects the overall economy. Hence, it seems reasonable to include the market risk premium in the model, but remove its effect on the stock portion of the portfolio.5
Adding the market risk premium and adjusting asset returns to reflect only the amount in excess of the risk-free rate, improved explained portfolio variance for all but the Max Return portfolio. This is encouraging because we worried that removing the risk-free rate might cause some information loss, and hence explanatory power, from raw returns.
We believe the improvement in explained variance for three of the portfolios is due mainly to an increase in the risk model’s power to explain bond returns. While we don’t show it, the \(R^{2}\) for bonds increased about seven percentage points by moving from the raw return to the market risk premium model. Recall that only the Max Return portfolio has a small exposure to bonds, while the others are above 20%. So maybe the market risk premium for equities does confer some information for bonds. Many companies are in both buckets after all.
How does this all compare to the original model using only F-F? On first blush, not so well. While the Max Sharpe portfolio saw a huge jump in explanatory power, the remaining portfolios saw modest changes or actual declines. Here’s the graph of that original output.
Of course, that original model was on a coincident time scale, so it does little to help us forecast impending risk. Clearly, the addition of macro variables helped the Max Sharpe portfolio because it had a very high weighting to real estate, whose variance in returns was explained best by the macro and macro plus F-F models. Nonetheless, we don’t think we should be discouraged by the results. Quite the opposite. That we were able to maintain the magnitude of variance explained on forward returns, which are notoriously difficult to forecast anyway, may commend the use of macro variables.
This isn’t time to take a victory lap, however. Many open questions about the model remain. First, while not a (insert risk factor commercial provider here) 40 factor model, with 11 factors our model isn’t exactly parsimonious. Second, we arbitrarily chose a look back/look forward combination we liked. There wasn’t anything rigorous to it and there’s no evidence, without out-of-sample validation, that the model would generalize well on unseen data. Third, even with all this data wrangling, we still haven’t found a model that explains even half of the variance of the portfolios. True, it’s better than zero, but that begs the question as to which matters more: behavior or opportunity cost. That is, while the simplistic risk measure volatility doesn’t tell us much about the future, it does tell us that returns will vary greatly. The risk factor models tell us that some of the future variance may be explained by some factors we’ve observed today, but the majority of the variance is still unknown.
Are you more comfortable knowing that you won’t know or knowing that you’ll know some, but not a lot and it will cost you time (and/or money) to get there? With that open question we’ll end this post. And now for the best part… the code!
# Built with R 4.0.3 and Python 3.8.3 #[R code] ## Load packages suppressPackageStartupMessages({ library(tidyverse) library(tidyquant) library(reticulate) }) # Allow variables in one python chunk to be used by other chunks. knitr::knit_engines$set(python = reticulate::eng_python) # [Python code] #') DIR = "C:/Users/user/image_folder" # Whatever directory you want to save the graphs in. This is the static folder in blogdown so we can grab png files instead running all of the code every time we want to update the Rmarkdown file. ## Create save figure function def save_fig_blog(fig_id, tight_layout=True, fig_extension="png", resolution=300): path = os.path.join(DIR, fig_id + "." + fig_extension) print("Saving figure", fig_id) if tight_layout: plt.tight_layout() plt.savefig(path, format=fig_extension, dip=resolution) ## Random portfolios for later # See other posts for Port_sim.calc_sim_lv() function np.random.seed(42) port1, wts1, sharpe1 = Port_sim.calc_sim_lv(df = df.iloc[1:60, 0:4], sims = 10000, cols=4) ## Download and wrangle macro variables import quandl quandl.ApiConfig.api_key = 'your_key' start_date = '1970-01-01' end_date = '2019-12-31' pmi = quandl.get("ISM/MAN_PMI", start_date=start_date, end_date=end_date) pmi = pmi.resample('M').last() # PMI is released on first of month. But most of the other data is on the last of the month start_date = '1970-01-01' end_date = '2019-12-31' indicators = ['ICSA', 'PERMIT','USSLIND', 'BOGMBASE', 'T10Y2Y', 'DAAA', 'DBAA'] fred_macros = {} for indicator in indicators: fred_macros[indicator] = dr.DataReader(indicator, 'fred', start_date, end_date) fred_macros[indicator] = fred_macros[indicator].resample('M').last() from functools import reduce risk_factors = pmi for indicator in indicators: risk_factors = pd.merge(risk_factors, fred_macros[indicator], how = 'left', left_index=True, right_index=True) risk_factors['CORP'] = risk_factors['DBAA'] - risk_factors['DAAA'] risk_factors = risk_factors.drop(['DAAA', 'DBAA'], axis=1) # Transform to sequential change macro_risk_chg = risk_factors.copy() macro_risk_chg.loc[:, ['PMI','ICSA', 'PERMIT', 'BOGMBASE']] = macro_risk_chg.loc[:, ['PMI','ICSA', 'PERMIT', 'BOGMBASE']].pct_change() macro_risk_chg.loc[:, ['USSLIND','T10Y2Y', 'CORP']] = macro_risk_chg.loc[:, ['USSLIND','T10Y2Y', 'CORP']].apply(lambda x: x*.01) ## Remove infinites for OLS macro_risk_chg = macro_risk_chg.replace([np.inf, -np.inf], 0.0) ## Spaghetti graph of macro variables ax = (macro_risk_chg.loc['1987':'1991', macro_risk_chg.columns[:-2].to_list()]*100).plot(figsize=(12,6), cmap = 'Blues') # macro_risk_chg.loc['1987':'1991', macro_risk_chg.columns[-2:].to_list()].plot(ax = ax) # Include only if want yield spreads too plt.legend(['PMI', 'Initial Claims', 'Housing', 'Leading indicators', 'Monetary base']) plt.ylabel('Percent change (%)') plt.title('Macroeconomic risk variables') plt.show() ## Create function to calculate R-squareds, with or without risk premiums, graph them, and do a whole bunch of other stuff that makes analysis and blog writing easier import statsmodels.api as sm ## Load data risk_factors = pd.read_pickle('risk_factors_2.pkl') df = pd.read_pickle('port_const.pkl') df.iloc[0,3] = 0.006 dep_df_all = df.iloc[:,:-1] ## Create RSQ function # Allows one to remove market risk premium from stock regression def rsq_func_prem(ind_df, dep_df, look_forward = None, risk_premium = False, period = 60, start_date=0, \ plot=True, asset_names = True, print_rsq = True, chart_title = None,\ y_lim = None, save_fig = False, fig_name = None): """ Assumes ind_df starts from the same date as dep_df. Dep_df has only as many columns as interested for modeling. """ xs = ind_df[0:start_date+period] assets = dep_df.columns.to_list() rsq = [] if look_forward: start = start_date + look_forward end = start_date + look_forward + period else: start = start_date end = start_date + period if risk_premium: mask = [x for x in factors.columns.to_list() if x != 'mkt-rfr'] for asset in assets: if asset == 'stock': X = sm.add_constant(xs.loc[:,mask]) else: X = sm.add_constant(xs) y = dep_df[asset][start:end].values mod = sm.OLS(y, X).fit().rsquared*100 rsq.append(mod) if print_rsq: print(f'R-squared for {asset} is {mod:0.03f}') else: X = sm.add_constant(xs) for asset in assets: y = dep_df[asset][start:end].values mod = sm.OLS(y, X).fit().rsquared*100 rsq.append(mod) if print_rsq: print(f'R-squared for {asset}label("$R^{2}$") if chart_title: plt.title(chart_title) else: plt.title("$R^{2}$ for Macro Risk Factor Model") plt.ylim(y_lim) if save_fig: save_fig_blog(fig_name) else: plt.tight_layout() plt.show() return rsq # Run r-squared function import statsmodels.api as sm cols = macro_risk_chg.columns[:-2].to_list() ind_df_1 = macro_risk_chg.loc['1987':'1991',cols] dep_df_1 = df.iloc[:60,:4] _ = rsq_func_prem(ind_df_1, dep_df_1, y_lim = None, save_fig = True, fig_name = 'macro_risk_r2_22') ## Create factor beta calculation function ## while allowing for market risk premium exposure def factor_beta_risk_premium_calc(ind_df, dep_df, risk_premium = False): """ Assumes only necessary columns in both ind_df and dep_df data frames. Set risk_premium to True if you include the market risk premium in ind_df and don't want to regress stock returns against i. """ xs = ind_df factor_names = [x.lower() for x in ind_df.columns.to_list()] assets = dep_df.columns.to_list() betas = pd.DataFrame(index=dep_df.columns) pvalues = pd.DataFrame(index=dep_df.columns) error = pd.DataFrame(index=dep_df.index) if risk_premium: mask = [x for x in ind_df.columns.to_list() if x != 'mkt-rfr'] # remove market risk premium from independent variables zero_val = np.where(ind_df.columns == 'mkt-rfr')[0][0] # identify index of market risk premium for asset in assets: if asset == 'stock': X = sm.add_constant(xs.loc[:,mask]) y = dep_df[asset].values result = sm.OLS(y, X).fit() # pad results for missing market risk premium results = np.array([x for x in result.params[1:zero_val+1]] + [0.0] + [x for x in result.params[zero_val+1:]]) for j in range(len(results)): # results and factor names have same length betas.loc[asset, factor_names[j]] = results[j] pvalues.loc[asset, factor_names[j]] = results[j] else: X = sm.add_constant(xs) y = dep_df[asset].values result = sm.OLS(y, X).fit() for j in range(1, len(result.params)): # result.params equals length of factor_names + 1 due to intercept so start at 1 betas.loc[asset, factor_names[j-1]] = result.params[j] pvalues.loc[asset, factor_names[j-1]] = result.pvalues[j] # Careful of error indentation: lopping through assets error.loc[:,asset] = (y - X.dot(result.params)) else: X = sm.add_constant(xs) for asset in assets: y = dep_df[asset].values result = sm.OLS(y, X).fit() for j in range(1, len(result.params)): betas.loc[asset, factor_names[j-1]] = result.params[j] pvalues.loc[asset, factor_names[j-1]] = result.pvalues[j] error.loc[:,asset] = (y - X.dot(result.params)) return betas, pvalues, error ## Create function to plot betas def betas_plot(beta_df, colors, legend_names, save_fig = False, fig_name=None): beta_df.plot(kind='bar', width = 0.75, color= colors, figsize=(12,6)) plt.legend(legend_names) plt.xticks([0,1,2,3], ['Stock', 'Bond', 'Gold', 'Real estate'], rotation=0) plt.ylabel(r'Factor $\beta$s') plt.title(r'Factor $\beta$s by asset class') if save_fig: save_fig_blog(fig_name) plt.show() ## Create function to calculate variance due to risk factors ## Create weight variables to calculate portfolio explained variance]] ## Calculate portfolio variance wt_list = [satis_wt, equal_wt, max_sharp_wt, max_ret_wt] port_exp=[] betas1, pvalues1, error1 = factor_beta_risk_premium_calc(ind_df_1, dep_df_1) for wt in wt_list: out = factor_port_var(betas1, ind_df_1, wt, error1) port_exp.append(out[0]/(out[0] + out[1])) port_exp = np.array(port_exp) ## Create function to plot portfolio explained variance def port_var_plot(port_exp, port_names=None, y_lim=None, save_fig=False, fig_name=None): if not port_names: port_names = ['Satisfactory', 'Naive', 'Max Sharpe', 'Max Return'] else: port_names = port_names plt.figure(figsize=(12,6)) plt.bar(port_names, port_exp*100, color='blue') for i in range(4): plt.annotate(str(round(port_exp[i]*100,1)) + '%', xy = (i-0.05, port_exp[i]*100+0.5)) plt.title('Original four portfolios variance explained by Macro risk factor model') plt.ylabel('Variance explained (%)') plt.ylim(y_lim) if save_fig: save_fig_blog(fig_name) plt.show() # Graph portfolio variance based on macro model port_var_plot(port_exp, y_lim = [0,22]) ## Run grid search based on look back normalization and look forward returns dep_df_all = df.iloc[:,:-1] # Grid search for best params ## Sort data scaled_sort = scale_for.sort_values(['Stocks','Bonds'], ascending=False).round(1).reset_index() # [R] ## Print table py$scaled_sort %>% as.data.frame(check.names=FALSE) %>% select(-index) %>% slice(1:5) %>% knitr::kable('html', caption = "Top 5 combinations by $R^{2}$ for stocks and bonds") ## Build model based on grid search model using 5-8 look back/look forward scale_5_8 = risk_factors.apply(lambda x: (x - x.rolling(5).mean())/x.rolling(5).std(ddof=1))['1987':] scale_5_8.replace([np.inf, -np.inf], 0.0, inplace=True) _ = rsq_func_prem(scale_5_8, dep_df_all, look_forward = 8, y_lim = [0,35],save_fig=False) betas_scale, _, error_scale = factor_beta_risk_premium_calc(scale_5_8['1987':'1991'],df.iloc[8:68,:-1]) beta_colors = ['darkblue', 'darkgrey', 'blue', 'grey','lightblue', 'lightgrey', 'black'] beta_names = ['PMI', 'Initial claims', 'Housing','Leading indicators', 'Monetary base', 'Treasury', 'Corp'] betas_plot(betas_scale, beta_colors, beta_names, save_fig=True, fig_name = 'factor_betas_5_8_22') ## Graph portfolio explained variance based on 5-8 grid search model_scale, scale_5_8['1987':'1991'], wt, error_scale) port_exp.append(out[0]/(out[0] + out[1])) port_exp = np.array(port_exp) port_var_plot(port_exp, y_lim=[0,30], save_fig=True, fig_name='four_port_var_exp_5_8_22') ## Load Fama-French Factors try: ff_mo = pd.read_pickle('ff_mo.pkl') print('Data loaded') except FileNotFoundError: print("File not found") print('Loading data yo...') ff_url = "" col_names = ['date', 'mkt-rfr', 'smb', 'hml', 'rfr'] ff = pd.read_csv(ff_url, skiprows=3, header=0, names = col_names) ff = ff.iloc[:1132,:].set_index('date', inplace=True) ff_mo = ff_mo*.01 # Convert percentages to decimal. IMPORTANT FOR RISK PREMIUM!! # Create total factors model total_factors = pd.merge(risk_factors, ff_mo, how='left', left_index=True, right_index=True) ## Grid search for best params # Prepare data dep_df_all = df.iloc[:,:-1] # Risk factor data keep = [x for x in total_factors.columns.to_list() if x not in ['mkt-rfr', 'rfr']] risk_factors_1 = total_factors.loc[:, keep]_1 scaled_sort = scale_for.sort_values(['Stocks', 'Bonds'], ascending=False).round(1).reset_index() [R] ## Print table from Python data frame py$scaled_sort %>% as.data.frame(check.names=FALSE) %>% select(-index) %>% slice(1:5) %>% knitr::kable('html', caption = "Top 5 combinations by $R^{2}$ for stocks and bonds of combined risk factor model") ## Build grid search model and graph portfolio variance scale_10_5 = risk_factors_1.apply(lambda x: (x - x.rolling(10).mean())/x.rolling(10).std(ddof=1))['1987':] scale_10_5.replace([np.inf, -np.inf], 0.0, inplace=True) betas_10_5, _, error_10_5 = factor_beta_risk_premium_calc(scale_10_5['1987':'1991'],df.iloc[5:65,:-1]), scale_10_5['1987':'1991'], wt, error_10_5) port_exp.append(out[0]/(out[0] + out[1])) port_exp = np.array(port_exp) port_var_plot(port_exp, y_lim=[0,40], save_fig=True, fig_name = 'four_port_var_exp_10_5_22') ## Run gird search for risk premium and market-risk premium factor model ### Grid search for best params ## Prepare data # Risk premium data dep_df_all = df.iloc[:,:-1].apply(lambda x: x - total_factors.loc['1987':'2019','rfr'].values) # Risk factor data risk_factors_3 = total_factors.loc[:, total_factors.columns.to_list()[:-1]] ## Create data frame scale_for = pd.DataFrame(np.c_[np.array([np.repeat(x,12) for x in range(3,13)]).flatten(),\ np.array([np.arange(1,13)]*10).flatten(),\ np.array([np.zeros(120)]*4).T],\ columns = ['Window', 'Forward', 'Stocks', 'Bonds', 'Gold', 'Real estate']) ## Iterate count = 0 for i in range(3, 13): risk_scale = risk_factors_3.apply(lambda x: (x - x.rolling(i).mean())/x.rolling(i).std(ddof=1))['1987':] risk_scale.replace([np.inf, -np.inf], 0.0, inplace=True) for j in range(1,13): out = rsq_func_prem(risk_scale, dep_df_all, risk_premium = True, look_forward = j, plot=False, print_rsq=False) scale_for.iloc[count,2:] = np.array(out) count+=1 scale_10_5 = risk_factors_3.apply(lambda x: (x - x.rolling(10).mean())/x.rolling(10).std(ddof=1))['1987':] scale_10_5.replace([np.inf, -np.inf], 0.0, inplace=True) _ = rsq_func_prem(scale_10_5, dep_df_all, look_forward = 5, risk_premium = True, y_lim = [0,60],save_fig=False) betas_10_5a, _, error_10_5a = factor_beta_risk_premium_calc(scale_10_5['1987':'1991'],dep_df_all.iloc[5:65,:], risk_premium=True)a, scale_10_5['1987':'1991'], wt, error_10_5a) port_exp.append(out[0]/(out[0] + out[1])) port_exp = np.array(port_exp) port_var_plot(port_exp, y_lim=[0,42], save_fig=True, fig_name='four_port_var_10_5_rp_22')
We apologize to our non-US readers. The US-bias is data-driven not philosophical. If any kind soul knows of easy to procure international data of sufficient length and frequency (i.e., more than 30 years and at least monthly), please let us know at content at optionstocksmachines.com. We’ll be happy to incorporate it into a post or three!↩︎
From the FRED perspicuous reader will note that there is some double counting since we also include housing permits in the model. However, when we used only the leading index, we didn’t find much to commend the model. Hopefully, the impact of permits in the leading index is modest.↩︎
Around 70% of the world’s population lives in developed countries and this is the main investing population. Of the more developed countries over 50% of the population is in the 25-64 years old range, which represents say 90% of investors. And if the most often recommended portfolio is some combination of 80/20 to 60/40 stocks/bonds, then 50-70% of portfolios will be overweight stocks.↩︎
We’re using the Wilshire 5000 total return index.↩︎
This is, of course, easy enough to do if we were just writing procedural code; writing functional code, with proper evaluators to ensure successful matrix multiplication is where it becomes a bit more gnarly!↩︎
Want to share your content on python-bloggers? click here. | https://python-bloggers.com/2021/01/more-factors-more-variance-explained/ | CC-MAIN-2021-10 | en | refinedweb |
Wrapper nodekit class. More...
#include <Inventor/nodekits/SoWrapperKit.h>
Wrapper nodekit class.
SoWrapperKit is derived from SoSeparatorKit. It adds the capability to wrap an arbitrary scene graph, (non-nodekit), within an SoSeparatorKit, so that it may be used along with other shape kits in a hierarchy. There are two additional parts included in SoWrapperKit: localTransform and contents .
The part contents is an SoSeparator node beneath which any arbitrary scene graph can be added. This is especially useful for importing scene graphs of unknown structure (non-nodekits) into nodekit format.
Since an SoWrapperKit is a class descended from SoSeparatorKit, it may be put into another SoSeparatorKit's childList .
(SoTransform) localTransform
This part is an SoTransform node that is used to affect the scene graph defined in contents part. This part is NULL by default and is automatically created if requested.
(SoSeparator) contents
This part is an SoSeparator node that contains a user-supplied scene graph. This scene graph can contain any nodes. This part is NULL by default and an SoSeparator is automatically created if the user asks the nodekit to build the part.
Extra Information for List Parts from Above Table
SoAppearanceKit, SoBaseKit, SoCameraKit, SoLightKit, SoNodeKit, SoNodeKitDetail, SoNodeKitListPart, SoNodeKitPath, SoNodekitCatalog, SoSceneKit, SoSeparatorKit, SoShapeKit
Constructor.
Returns the SoNodekitCatalog for this class.
Reimplemented from SoSeparatorKit.
Returns the type identifier for this class.
Reimplemented from SoSeparatorKit.
Returns the SoNodekitCatalog for this instance.
Reimplemented from SoSeparatorKit.
Returns the type identifier for this specific instance.
Reimplemented from SoSeparatorKit. | https://developer.openinventor.com/refmans/9.9/RefManCpp/class_so_wrapper_kit.html | CC-MAIN-2021-10 | en | refinedweb |
Learn Python - Python tutorial - python while loop - Python examples - Python programs.
python - Sample - python code :
# prints Hello Wikitechy 3 Times count = 0 while (count < 3): count = count+1 print("Hello Wikitechy")
python programming - Output :
Hello Wikitechy Hello Wikitechy Hello Wikitechy.
python - Sample - python code :
# Iterating over a list print("List Iteration") l = ["wikitechy", "best e-learning", "website"] for i in l: print(i) # Iterating over a tuple (immutable) print("\nTuple Iteration") t = ("wikitechy", "best e-learning", "website") for i in t: print(i) # Iterating over a String print("\nString Iteration") s = "Wikitechy" for i in s : print(i) # Iterating over dictionary print("\nDictionary Iteration") d = dict() d['xyz'] = 123 d['abc'] = 345 for i in d : print("%s %d" %(i, d[i]))
python programming - Output :
List Iteration wikitechy best e-learning website Tuple Iteration wikitechy best e-learning website String Iteration W i k i t e c h y Dictionary Iteration xyz 123 abc 345
We can use for in loop for user defined iterators.:
Syntax:
while expression: while expression: statement(s) statement(s)
A final note on loop nesting is that we can put any type of loop inside of any other type of loop. For example a for loop can be inside a while loop or vice versa.
python - Sample - python code :
from __future__ import print_function for i in range(1, 5): for j in range(i): print(i, end=' ') print()
python programming -.
python - Sample - python code :
# Prints all letters except 'e' and 's' for letter in 'wikitechy': if letter == 'i' or letter == 'y': continue print 'Current Letter :', letter var = 10
python programming - Output :
Current Letter : w Current Letter : k Current Letter : t Current Letter : e Current Letter : c Current Letter : h
Break Statement
It brings control out of the loop
python - Sample - python code :
for letter in 'wikitechy': # break the loop as soon it sees 'i' # or 'c' if letter == 'i' or letter == 'c': break print 'Current Letter :', letter
python programming - Output :
Current Letter : i
Pass Statement
We use pass statement to write empty loops. Pass is also used for empty control statement, function and classes.
python - Sample - python code :
# An empty loop for letter in 'wikitechy': pass print 'Last Letter :', letter
python programming - Output :
Last Letter : c | https://www.wikitechy.com/tutorials/python/python-while | CC-MAIN-2021-10 | en | refinedweb |
Closed Bug 722591 Opened 9 years ago Closed 9 years ago
Longtap on page title (not awesomescreen) should show an editbox context menu
Categories
(Firefox for Android Graveyard :: General, defect, P3)
Tracking
(firefox15 verified, firefox16 verified, firefox17 verified, fennec+)
Firefox 16
People
(Reporter: wesj, Assigned: wesj)
References
Details
Attachments
(2 files, 6 obsolete files)
I've been pasting a url in a lot lately because I constantly restart. Pasting the url requires tapping urlbar, waiting for the awesomescreen to show, and then long tapping and pasting into the awesomescreen textbox. We should show the same url (and maybe also open the awesomescreen?) when you long tap on the page title.
What about the long tap menu on the title including "paste and go" -- maybe "copy URL" and "copy title"? Just thinking out loud here...
Assignee: nobody → wjohnston
tracking-fennec: --- → +
Priority: -- → P3
As part of my 30min/day of paper cut fixes, a WIP to make this happen. Works. Needs string and id cleanup, and a test of course.
I needed a little break, so this adds "Paste & Go", "Paste", "Share", "Copy Location", "Copy Title", and "Add to Home Screen" to the titlebar context menu. I always feel like this context menu code is a mess, so hoping margaret can tell me the right, clean way to do it. Now back to real work...
Attachment #601138 - Attachment is obsolete: true
Attachment #627347 - Flags: review?(margaret.leibovic)
In the name of keeping things simple, I think we should probably get rid of "Copy Title" which isn't ever very useful.
Updated with the missing file.
Attachment #627347 - Attachment is obsolete: true
Attachment #627347 - Flags: review?(margaret.leibovic)
Attachment #628425 - Flags: review?(margaret.leibovic)
Comment on attachment 628425 [details] [diff] [review] Patch v2.5 Review of attachment 628425 [details] [diff] [review]: ----------------------------------------------------------------- Holding off review until there's a new (un-bitrotted) version of the patch. Here are some comments to start you off, though :) I promise to get back to you sooner next time around! ::: mobile/android/base/BrowserToolbar.java @@ +119,5 @@ > + Tab tab = Tabs.getInstance().getSelectedTab(); > + if (tab != null) { > + String url = tab.getURL(); > + if (url == null) { > + menu.findItem(R.id.copyurl).setVisible(false); We probably also want to hide any tab-dependant menu items when tab is null (like share, etc.). ::: mobile/android/base/GeckoApp.java @@ +507,5 @@ > } > > + private boolean shareCurrentUrl() { > + Tab tab = Tabs.getInstance().getSelectedTab(); > + if (tab != null) { Nit: I prefer just returning if tab is null, to avoid indenting the rest of the body of the method. @@ +515,5 @@ > + > + GeckoAppShell.openUriExternal(url, "text/plain", "", "", > + Intent.ACTION_SEND, tab.getTitle()); > + } > + return true; I know you just copied this, but why are we returning true here even if the tab is null, considering we return false if the url is null? Reading the docs, it looks like we always just want to return true, since that's just used to indicate we're handing the menu item: I propose we make shareCurrentUrl just return void, then stick the return true back below it in onOptionsItemSelected. @@ +2898,5 @@ > public boolean linkerExtract() { > return false; > } > + > + public boolean onContextItemSelected(MenuItem item) { I think you want an @Override above this. @@ +2914,5 @@ > + if (text != null && !TextUtils.isEmpty(text)) { > + showAwesomebar(AwesomeBar.Type.ADD, text); > + return true; > + } > + break; I don't like that there's a mix of breaks and returns in here. I think we should replace the breaks with returns. @@ +2917,5 @@ > + } > + break; > + } > + case R.id.share: { > + return shareCurrentUrl(); This would need to change if you change shareCurrentUrl like I suggested above. @@ +2944,5 @@ > + } > + return true; > + } > + } > + return false; My same questions about returning true/false apply to this method: It seems to me like we should always be returning true if the user taps on one of the menuitems that we created. I'm not sure what returning false would mean.
Attachment #628425 - Flags: review?(margaret.leibovic) → feedback+
Unbitrotted and feedback adjusted. Thanks!
Attachment #628425 - Attachment is obsolete: true
Attachment #630309 - Flags: review?(margaret.leibovic)
Comment on attachment 630309 [details] [diff] [review] Unbitrotted Patch I'm giving this an r- mainly because I think we should get UX feedback on what Paste and Paste&Go should do w.r.t. new/current tabs. >diff --git a/mobile/android/base/GeckoApp.java b/mobile/android/base/GeckoApp.java >@@ -652,16 +644,30 @@ abstract public class GeckoApp >+ private void shareCurrentUrl() { >+ Tab tab = Tabs.getInstance().getSelectedTab(); >+ if (tab == null) >+ return; >+ >+ String url = tab.getURL(); >+ if (url == null) >+ return; >+ >+ GeckoAppShell.openUriExternal(url, "text/plain", "", "", >+ Intent.ACTION_SEND, tab.getTitle()); >+ return; You don't need this return at the end of the method :) >@@ -2759,29 +2765,35 @@ abstract public class GeckoApp > public boolean showAwesomebar(AwesomeBar.Type aType) { >+ return showAwesomebar(aType, null); >+ } >+ >+ public boolean showAwesomebar(AwesomeBar.Type aType, String aUrl) { > Intent intent = new Intent(getBaseContext(), AwesomeBar.class); > intent.addFlags(Intent.FLAG_ACTIVITY_NO_ANIMATION | Intent.FLAG_ACTIVITY_NO_HISTORY); > intent.putExtra(AwesomeBar.TYPE_KEY, aType.name()); > >- if (aType != AwesomeBar.Type.ADD) { >+ if (aUrl != null || aType != AwesomeBar.Type.ADD) { Can we change the second part of this to |if aType == AwesomeBar.Type.EDIT|? That's the only other case, and it makes it clearer what's going on here. > // if we're not adding a new tab, show the old url >- Tab tab = Tabs.getInstance().getSelectedTab(); >- if (tab != null) { >- String url = tab.getURL(); >- if (url != null) { >- intent.putExtra(AwesomeBar.CURRENT_URL_KEY, url); >+ if (aUrl == null) { >+ Tab tab = Tabs.getInstance().getSelectedTab(); >+ if (tab != null) { >+ aUrl = tab.getURL(); > } > } >+ if (aUrl != null) { >+ intent.putExtra(AwesomeBar.CURRENT_URL_KEY, aUrl); >+ } > } This logic feels too messy, especially since your new use case only needs to handle a call with Type.ADD. Maybe we should only do something with the extra aUrl param in the ADD case? I just think we need to be less clever here, and tease apart the different cases (and some comments explaining the logic would be good). I also think we should get some UX feedback here on whether or not the result of "Paste" would be expected to open a new tab or in the current tab, especially since Paste&Go as you have it overrides the current tab. >@@ -3078,9 +3090,56 @@ abstract public class GeckoApp >+ @Override >+ public boolean onContextItemSelected(MenuItem item) { >+ switch(item.getItemId()) { ... >+ } >+ return true; I think the correct thing to do is to return false at the end of this method in case there's somehow a context menu item that we didn't put there, but return true inside each of the case statements when we know for sure we handled the menu item.
Attachment #630309 - Flags: review?(margaret.leibovic) → review-
Ian, can you help us out here? Do you think Paste/Paste&Go should just modify the url of the currently selected tab? Because that's what I think :) (The other option is for one or both of them to open the url in a new tab)
I'll wear you down eventually! I agree, these should open in the same tab. Not sure why I did that the way I originally did. Also cleaned up the showAwesomebar logic a bit.
Attachment #630379 - Flags: review?(margaret.leibovic)
This converts AwesomeBar.Type to AwesomeBar.Target with values NEW_TAB and CURRENT_TAB. Helps me not make mistakes.
Attachment #630309 - Attachment is obsolete: true
Attachment #630380 - Flags: review?(margaret.leibovic)
Margaret and Wes, I also agree that the pasted URL should affect the current tab, so it's unanimous!
Comment on attachment 630379 [details] [diff] [review] Patch Review of attachment 630379 [details] [diff] [review]: ----------------------------------------------------------------- Looking good! Just one last comment I found... ::: mobile/android/base/BrowserToolbar.java @@ +119,5 @@ > + Tab tab = Tabs.getInstance().getSelectedTab(); > + if (tab != null) { > + String url = tab.getURL(); > + if (url == null) { > + menu.findItem(R.id.copyurl).setVisible(false); I just realized that share and add to home screen won't work out the url is null, so we can also hide those here.
Attachment #630379 - Flags: review?(margaret.leibovic) → review+
Comment on attachment 630380 [details] [diff] [review] Cleanup Types Review of attachment 630380 [details] [diff] [review]: ----------------------------------------------------------------- ::: mobile/android/base/AwesomeBar.java @@ +63,2 @@ > > private String mType; I think we should also rename mType to mTarget. Kill "type"!
Attachment #630380 - Flags: review?(margaret.leibovic) → review- (Merged by Ed Morley)
Status: NEW → RESOLVED
Closed: 9 years ago
Resolution: --- → FIXED
Target Milestone: --- → Firefox 16
This introduced a NPE crash in bug 763167.
Uhmm... I ignored your comment... :(
Oh wait, I did make this change on checkin. Here's the patch that went. I assume the r+ was from irc.... I hope. I'm looking at moving this to Aurora because I have another aurora+ patch that depends on it. I'm gonna unbit that around this though. I'll nom it anyway to protect other patches that might depend on this.
Attachment #630380 - Attachment is obsolete: true
Attachment #637291 - Attachment is obsolete: true
Attachment #637297 - Flags: review+
Comment on attachment 637297 [details] [diff] [review] Uploaded Patch [Approval Request Comment] Bug caused by (feature/regressing bug #): Initial Fennec work had this. User impact if declined: None. This is purely to make development a little easier by using better variable names. The point is to save us some potential conflicts with other patches going to Aurora. Testing completed (on m-c, etc.): Several months ago. Risk to taking this patch (and alternatives if risky): Very low risk. No changes. String or UUID changes made by this patch: None.
Attachment #637297 - Flags: approval-mozilla-aurora?
Verified fixed on: Htc Desire Z (2.3.3) Using: Nightly Fennec 16.0a1 (2012-07-09) The patch has not been added to aurora yet, although it was approved. Waiting for the patch to land in Aurora.
status-firefox16: --- → verified
Whoops. Thanks:
Status: RESOLVED → VERIFIED
status-firefox17: --- → verified
Product: Firefox for Android → Firefox for Android Graveyard | https://bugzilla.mozilla.org/show_bug.cgi?id=722591 | CC-MAIN-2021-10 | en | refinedweb |
SIGPAUSE(3B) SIGPAUSE(3B)
sigpause - atomically release blocked signals and wait for interrupt
(4.3BSD)
#include <signal.h>
int sigpause(int mask);
mask = sigmask(int signpause assigns mask to the set of masked signals and then waits for a
signal to arrive; upon return the original set of masked signals is
restored after executing the handler(s) (if any) installed for the
awakening signal(s). mask is usually 0 to indicate that no signals are
now to be blocked. The macro sigmask is provided to construct the mask
for a given signal number. Sigpause always terminates by being
interrupted, returning -1 with the global integer errno set to EINTR.
In normal usage, a signal is blocked using sigblock(3B), to begin a
critical section, variables modified on the occurrence of the signal are
examined to determine that there is no work to be done, and the process
pauses awaiting work by using sigpause with the mask returned by
sigblock.
sigblock(3B), sigvec(3B), signal(5).
CAVEATS (IRIX)
Because 4.3BSD and System V both have sigpause system calls, programs
using 4.3BSD's version are actually executing BSDsigpause. This is
transparent to the programmer except when attempting to set breakpoints
in dbx; the breakpoint must be set at BSDsigpause.
WARNING (IRIX)
The 4.3BSD and System V signal facilities have different semantics.
Using both facilities in the same program is strongly discouraged and
will result in unpredictable behavior.
PPPPaaaaggggeeee 1111 | https://nixdoc.net/man-pages/IRIX/man3B/sigpause.3B.html | CC-MAIN-2021-10 | en | refinedweb |
public class FirebaseCustomLocalModel extends FirebaseLocalModel
This class is deprecated.
For more information refer to the custom model implementation instructions.
Describes a local model created from local or asset files.
Defines the model's location (either absolute file path in local directory, or subpath in the app's asset directory) and the model name.
Nested Class Summary
Inherited Method Summary
From class com.google.firebase.ml.common.modeldownload.FirebaseLocalModel
From class java.lang.Object | https://developers.google.cn/android/reference/com/google/firebase/ml/custom/FirebaseCustomLocalModel | CC-MAIN-2021-10 | en | refinedweb |
Calculated Fields
In this article you will find out how to extend RadPivotGrid's generated report by adding Calculated Fields.
Calculated Fields Description
If your data analysis requires results that are not available using just the data source fields and RadPivotGrid's built-in calculations, you can insert a calculated field that uses a custom formula to derive the results you need. A calculated field is a new data field in which the values are the result of a custom calculation formula. You can display the calculated field along with another data field or on its own. A calculated field is really a custom summary calculation, so in almost all cases, the calculated field references one or more fields in the source data.
Define Calculated Fieldider you have to add it to its CalculatedFields collection. So first you have to create a concrete class that implements the abstract class CalculatedField. This requires the implementation of two methods - CalculatedValue and RequiredFields. CalculateValue is the method in which you have to define your calculation formula and create a new AggregateValue that will be shown in RadPivotGrid. In common scenario a Calculated Field reference one or more fields in the data source. The RequiredFields method should return IEnumerable of the required fields. That's why we have added a new class called RequiredField. It's purpose is to describe a field, required for a calculated field. RequiredField can be created for property from the data object or for another calculated field. Name property of CalculatedField class identifies its unique name that will be shown in the UI.
The first task is to decide what is the calculation formula that you want to use. For example, you can show the commission that will be paid to all salespeople. Commission will be paid only to those who have more sold for more than $15 000. The price of the sold items is kept by the ExtendedPrice property from the source. So the new class will look like this:
public class CommissionCalculatedField : CalculatedField { private RequiredField extendPriceField; public CommissionCalculatedField() { this.Name = "Commission"; this.extendPriceField = RequiredField.ForProperty("ExtendedPrice"); } protected override IEnumerable<RequiredField> RequiredFields() { yield return this.extendPriceField; } protected override AggregateValue CalculateValue(IAggregateValues aggregateValues) { var aggregateValue = aggregateValues.GetAggregateValue(this.extendPriceField); if (aggregateValue.IsError()) { return aggregateValue; } double extendedPrice = aggregateValue.ConvertOrDefault<double>(); if (extendedPrice > 15000) { return new DoubleAggregateValue(extendedPrice * 0.1); } return null; } }
Public Class CommissionCalculatedField Inherits CalculatedField Private extendPriceField As RequiredField Public Sub New() Me.Name = "Commission" Me.extendPriceField = RequiredField.ForProperty("ExtendedPrice") End Sub Protected Overrides Iterator Function RequiredFields() As IEnumerable(Of RequiredField) Yield Me.extendPriceField End Function Protected Overrides Function CalculateValue(ByVal aggregateValues As IAggregateValues) As AggregateValue Dim aggregateValue = aggregateValues.GetAggregateValue(Me.extendPriceField) If aggregateValue.IsError() Then Return aggregateValue End If Dim extendedPrice As Double = aggregateValue.ConvertOrDefault(Of Double)() If extendedPrice > 15000 Then Return New DoubleAggregateValue(extendedPrice * 0.1) End If Return Nothing End Function
Now it is time to add a new instance of this class to the CalculatedFields collection of LocalDataSourceProvider:
<pivot:LocalDataSourceProvider.CalculatedFields> <local:CommissionCalculatedField </pivot:LocalDataSourceProvider.CalculatedFields>
var calculatedField = new CommissionCalculatedField(); calculatedField.Name = "Commission"; dataProvider.CalculatedFields.Add(calculatedField);
Dim calculatedField = New CommissionCalculatedField() calculatedField.Name = "Commission" dataProvider.CalculatedFields.Add(calculatedField)
If you add caculated fields in code behind, you have to set the ItemsSource of LocalDataSourceProv LocalDataSourceProvider's AggregateDescriptions collection:
<pivot:LocalDataSourceProvider.AggregateDescriptions> <pivot:CalculatedAggregateDescription </pivot:LocalDataSourceProvider.AggregateDescriptions>
var calculatedAggregate = new CalculatedAggregateDescription(); calculatedAggregate.CalculatedFieldName = "Commission"; dataProvider.AggregateDescriptions.Add(calculatedAggregate);
Dim calculatedAggregate = New CalculatedAggregateDescription() calculatedAggregate.CalculatedFieldName = "Commission" dataProvider.AggregateDescriptions.Add(calculatedAggregate)
The result will look like this:
| https://docs.telerik.com/devtools/silverlight/controls/radpivotgrid/features/localdatasourceprovider/local-calc-fields | CC-MAIN-2021-10 | en | refinedweb |
Code Review: SharpArchitecture.MultiTenant
It has been suggested that I’ll look at a more modern implementation of SharpArchitecture, and I was directed toward the MultiTenant project.
The first thing to notice is the number of projects. It is actually hard to keep the number of projects down, as I well know, but this has several strange choices.
I am not really sure what is the point in separating the controllers into a separate assembly, or why we have a separate project for the ApplicationServices.
I am not the only one thinking so, I think:
Then there is the Core project:
Personally, I wouldn’t create a project for just two files, but I can live with that. I don’t like attributes like DomainSignature. It is hard for me to really say what, except that I think that they encourage a way of thinking that puts the model in the Center of All Things. I am usually much more interested in what something is doing than how it is shaped.
The data project is mostly concerned with setting up NHibernate via Fluent NHibernate.
Next up is the Framework project. And there we run into the following marker interfaces. I really don’t like marker interfaces, and having those here doesn’t seem to be adding anything important to the application.
It seems that there is a lot going on simply to try to get a fallback to a non tenant situation, but the problem here is that it is usually much better to be explicit about those sort of things. You have the CentralSession, and you have the TenantSession, and you are working with each in a different manner. It makes the infrastructure easier to manage and usually result in code that is clearer to follow.
So far, it has all been pretty much uninteresting, I would strongly encourage merging the solution into just two projects, the web & the tests projects, but that is about it.
Now we move into the fancy bits, the controllers project. And there we find the following piece of code:
public class TenantListQuery : NHibernateQuery, ITenantListQuery
{
public IPagination<TenantViewModel> GetPagedList(int pageIndex, int pageSize)
{
var query = Session.QueryOver<Tenant>()
.OrderBy(customer => customer.Name).Asc;
var countQuery = query.ToRowCountQuery();
var totalCount = countQuery.FutureValue<int>();
var firstResult = (pageIndex - 1) * pageSize;
TenantViewModel viewModel = null;
var viewModels = query.SelectList(list => list
.Select(mission => mission.Id).WithAlias(() => viewModel.Id)
.Select(mission => mission.Name).WithAlias(() => viewModel.Name)
.Select(mission => mission.Domain).WithAlias(() => viewModel.Domain)
.Select(mission => mission.ConnectionString).WithAlias(() => viewModel.ConnectionString))
.TransformUsing(Transformers.AliasToBean(typeof(TenantViewModel)))
.Skip(firstResult)
.Take(pageSize)
.Future<TenantViewModel>();
return new CustomPagination<TenantViewModel>(viewModels, pageIndex, pageSize, totalCount.Value);
}
}
I quite like this code. It is explicit about what it is doing. There is a good reason to hide this sort of thing behind a class, because while it is easy to read, it is also a lot of detailed code that should be abstracted. I like the use of futures to reduce the number of queries, and that we have explicit paging here. I also like the projection directly into the view model.
What I don’t like is that I really don’t understand how the Session instance is being selected. Oh, I understand how MultiTenantSessionFactoryKeyProvider is working, and that we get the central database because we aren’t using a tenanted entity, but it still seems too much magic here I would rather a CentralSession instead.
Another thin that I liked was the code structure:
All of the code is grouped by feature in a very nice fashion.
My main peeve with the application is that this is basically it. We are talking about an application that is basically two CRUD pages, nothing more. Yes, it is a sample app to show something very specific, but I would have liked to see some more meat there to look at.
Multi tenancy is a hard problem, and this application spend quite a bit of time doing what is essentially connection string management.
(Note: Opinions expressed in this article and its replies are the opinions of their respective authors and not those of DZone, Inc.) | http://www.dzone.com/news/code-review | crawl-003 | en | refinedweb |
#include <BCP_lp_user.hpp>
Inheritance diagram for BCP_lp_user:
In that derived class the user can store data to be used in the methods she overrides. Also that is the object the user must return in the USER_initialize::l 72 of file BCP_lp_user.hpp.
Definition at line 144 of file BCP_lp_user.hpp.
Being virtual, the destructor invokes the destructor for the real type of the object being deleted.
Definition at line 147 of file BCP_lp_user.hpp.
Set the pointer.
Definition at line 90 of file BCP_lp_user.hpp.
Get the pointer.
Definition at line 92 of file BCP_lp_user.hpp.
Definition at line 95 of file BCP_lp_user.hpp.
References babSolver_.
Definition at line 96 of file BCP_lp_user.hpp.
References babSolver_.
Return what is the best known upper bound (might be BCP_DBL_MAX).
Return the phase the algorithm is in.
Return the level of the search tree node being processed.
Return the internal index of the search tree node being processed.
Return the iteration count within the search tree node being processed.
Return a pointer to the BCP_user_data structure the user (may have) stored in this node.
Select all nonzero entries.
Those are considered nonzero that have absolute value greater than
etol.
Select all zero entries.
Those are considered zero that have absolute value less than
etol.
Select all positive entries.
Those are considered positive that have value greater than
etol.
Select all fractional entries.
Those are considered fractional that are further than
etol away from any integer value. in BB_lp, MCF1_lp, MCF2_lp, and MCF3_lp. in BB_lp, MCF1_lp, MCF2_lp, and MCF3_lp. in BB_lp, MCF1_lp, MCF2_lp, and MCF3_lp.
Modify parameters of the LP solver before optimization.
This method provides an opportunity for the user to change parameters of the LP solver before optimization in the LP solver starts. The second argument indicates whether the optimization is a "regular" optimization or it will take place in strong branching. Default: empty method.
Process the result of an iteration.
This includes:
The default behavior is to do nothing and invoke the individual methods one-by-one. in MCF1_lp, MCF2_lp, and MCF3_lp. in BB_lp, MCF1_lp, MCF2_lp, and MCF3_lp.
Test whether all variables are 0/1.
Note that this method assumes that all variables are binary, i.e., their original lower/upper bounds are 0/1.
Test whether all variables are integer.
Note that this method assumes that all variables are integer.
Test whether the variables specified as integers are really integer.
Pack a MIP feasible solution into a buffer.
The solution will be unpacked in the Tree Manager by the BCP_tm_user::unpack_feasible_solution() method.
Default: The default implementation assumes that
sol is a BCP_solution_generic object (containing variables at nonzero level) and packs it.
Pack the information necessary for cut generation into the buffer.
Note that the name of the method is pack_primal_solution because most likely that (or some part of that) will be needed for cut generation. However, if the user overrides the method she is free to pack anything (of course she'll have to unpack it in CG).
This information will be sent to the Cut Generator (and possibly to the Cut Pool) where the user has to unpack it. If the user uses the built-in method here, then the built-in method will be used in the Cut Generator as well.
Default: The content of the message depends on the value of the
PrimalSolForCG parameter in BCP_lp_par. By default the variables at nonzero level are packed.
Pack the information necessary for variable generation into the buffer.
Note that the name of the method is pack_dual_solution because most likely that (or some part of that) will be needed for variable generation. However, if the user overrides the method she is free to pack anything (of course she'll have to unpack it in CG).
This information will be sent to the Variable Generator (and possibly to the Variable Pool) where the user has to unpack it. If the user uses the built-in method here, then the built-in method will be used in the Variable Generator as well.
Default: The content of the message depends on the value of the
DualSolForVG parameter in BCP_lp_par. By default the full dual solution is packed.
Display the result of most recent LP optimization.
This method is invoked every time an LP relaxation is optimized and the user can display (or not display) it.
Note that this method is invoked only if
final_lp_solution is true (i.e., no cuts/variables were found) and the
LpVerb_FinalRelaxedSolution parameter of BCP_lp_par is set to true (or alternatively,
final_lp_solution is false and
LpVerb_RelaxedSolution is true).
Default: display the solution if the appropriate verbosity code entry is set.
Return the index of the indexed variable following
prev_index.
Return -1 if there are no more indexed variables. If
prev_index is -1 then return the index of the first indexed variable.
Default: Return -1.
Create the variable corresponding to the given
index.
The routine should return a pointer to a newly created indexed variable and return the corresponding column in
col.
Default: throw an exception.
Restoring feasibility.
This method is invoked before fathoming a search tree node that has been found infeasible and the variable pricing did not generate any new variables.
If the MaintainIndexedVarPricingList is set to true then BCP will take care of going through the indexed variables to see if any will restore feasibility and the user has to check only the algorithmic variables. Otherwise the user has to check all variables here.
Convert (and possibly lift) a set of cuts into corresponding rows for the current LP relaxation.
Converting means computing for each cut the coefficients corresponding to each variable and creating BCP_row objects that can be added to the formulation.
This method has different purposes depending on the value of the last argument. If multiple expansion is not allowed then the user must generate a unique row for each cut. This unique row must always be the same for any given cut. This kind of operation is needed so that an LP relaxation can be exactly recreated.
On the other hand if multiple expansion is allowed then the user has (almost) free reign over what she returns. She can delete some of the
cuts or append new ones (e.g., lifted ones) to the end. The result of the LP relaxation and the origin of the cuts are there to help her to make a decision about what to do. For example, she might want to lift cuts coming from the Cut Generator, but not those coming from the Cut Pool. The only requirement is that when this method returns the number of cuts and rows must be the same and the i-th row must be the unique row corresponding to the i-th cut. in MCF1_lp, MCF2_lp, and MCF3_lp.
Generate cuts within the LP process.
Sometimes too much information would need to be transmitted for cut generation (e.g., the full tableau for Gomory cuts) or the cut generation is so fast that transmitting the info would take longer than generating the cuts. In such cases it might better to generate the cuts locally. This routine provides the opportunity.
Default: empty for now. To be interfaced to Cgl. in MCF1_lp, MCF2_lp, and MCF3_lp.
Compare two generated cuts.
Cuts are generated in different iterations, they come from the Cut Pool, etc. There is a very real possibility that the LP process receives several cuts that are either identical or one of them is better then another (cuts off everything the other cuts off). This routine is used to decide which one to keep if not both.
Default: Return
BCP_DifferentObjs.
Compare two generated variables.
Variables are generated in different iterations, they come from the Variable Pool, etc. There is a very real possibility that the LP process receives several variables that are either identical or one of them is better then another (e.g., almost identical but has much lower reduced cost). This routine is used to decide which one to keep if not both.
Default: Return
BCP_DifferentObjs.
Reduced cost fixing.
This is not exactly a helper function, but the user might want to invoke it... in BB_lp, MCF1_lp, MCF2_lp, and MCF3_lp.
This helper method creates branching variable candidates and appends them to
cans.
The indices (in the current formulation) of the variables from which candidates should be created are listed in
select_pos.
Decide which branching object is preferred for branching.
Based on the member fields of the two presolved candidate branching objects decide which one should be preferred for really branching on it. Possible return values are:
BCP_OldPresolvedIsBetter,
BCP_NewPresolvedIsBetter and
BCP_NewPresolvedIsBetter_BranchOnIt. This last value (besides specifying which candidate is preferred) also indicates that no further candidates should be examined, branching should be done on this candidate.
Default: The behavior of this method is governed by the
BranchingObjectComparison parameter in BCP_lp_par.
Decide what to do with the children of the selected branching object.
Fill out the
_child_action field in
best. This will specify for every child what to do with it. Possible values for each individual child are
BCP_PruneChild,
BCP_ReturnChild and
BCP_KeepChild. There can be at most child with this last action specified. It means that in case of diving this child will be processed by this LP process as the next search tree node.
Default: Every action is
BCP_ReturnChild. However, if BCP dives then one child will be mark with
BCP_KeepChild. The decision which child to keep is based on the
ChildPreference parameter in BCP_lp_par. Also, if a child has a presolved lower bound that is higher than the current upper bound then that child is mark as
BCP_FathomChild.
THINK*: Should those children be sent back for processing in the next phase?
For each child create a user data object and put it into the appropriate entry in
best->user_data().
When this function is called the
best->user_data() vector is already the right size and is filled will 0 pointers. The second argument is usefule if strong branching was done. It is the index of the branching candidate that was selected for branching (the one that's the source of
best.
Reimplemented in BB_lp, and MCF3_lp.
Deprecated version of the previos method (it does not pass the index of the selected branching candidate).
Selectively purge the list of slack cuts.
When a cut becomes ineffective and is eventually purged from the LP formulation it is moved into
slack_pool. The user might consider cuts might later for branching. This function enables the user to purge any cut from the slack pool (those she wouldn't consider anyway). Of course, the user is not restricted to these cuts when branching, this is only there to help to collect slack cuts. The user should put the indices of the cuts to be purged into the provided vector.
Default: Purges the slack cut pool according to the
SlackCutDiscardingStrategy rule in BCP_lp_par (purge everything before every iteration or before a new search tree node).
Definition at line 78 of file BCP_lp_user.hpp.
Definition at line 79 of file BCP_lp_user.hpp.
Referenced by getLpProblemPointer(), and setLpProblemPointer().
Definition at line 80 of file BCP_lp_user.hpp.
Referenced by getOsiBabSolver(), and setOsiBabSolver(). | http://www.coin-or.org/Doxygen/CoinAll/class_b_c_p__lp__user.html | crawl-003 | en | refinedweb |
There’s a new video up on which aims to help developers pick between ASP.NET WebForms and ASP.NET MVC. The video boils down to 5 benefits per technology which Microsoft thinks you should consider.
Let’s go over the points, shall we? First, ASP.NET WebForms:
1 – Familiar control and event base programming model
The claim here is that the ASP.NET model is comfortable for WinForm programmers (thankfully this unbiased analysis left out who it’s more familiar for). This is largely accurate, but disingenuous. The differences between web and desktop cannot be overstated nor can one overstate how bad ASP.NET’s (or any other framework) is at hiding the difference. “Familiar” is probably the right word to use so long as you recognize that, in this case, at best it means: superficial; at worst: a serious pain in the ass. Your knowledge of building a VB6 app will allow you to write a “Hello World” web application – great.
Familiarity can be a liability when it tries to force a square peg into a round hole.
It also largely relies on your inability (or unwillingness) to learn. Today, next month or even next year may not be the right time for you to learn something new – that’s fine. Eventually, sticking with what you know, only because you know it, will kill your career and possibility part of your spirit.
2 – Controls encapsulate HTML, JS and CSS
It’s true that in ASP.NET WebForms controls can, and frequently do, encapsulate HTML, JS and CSS. How this ads “value” is beyond me. You can’t, and shouldn’t, be trying to build website without a solid command of HTML, JS and CSS. Whatever programming language and framework you use, the ultimate output of any website is HTML, CSS and JavaScript. Your server code essentially generates a stream of characters, which a browser loads and renders. To suggest, or think, that generating HTML, CSS or JavaScript in C# has any advantage is insane. It’ll be more complicated to learn, do and maintain – and the end result will be inferior. Its like saying we should write C# in VB.NET; or drive cars by bolting planes to the roof and getting in the cockpit.
3 – Rich UI controls included – datagrids, charts, AJAX.
Point 3 is a different perspective on point 2, which is a different way of saying point 1. However, it is the most interesting and important perspective. Fancy tables and charts, as well as client-side behavior, shouldn’t be a server-side concern. This is fundamental to what we all know about good and bad design. Classic ASP was a mess because it intermingled presentation code with server side code. The value of WebForms is that presentation logic is now a server side concern. Do you really believe this? Would you consider generating your HTML from a stored procedure?
The claim also implies that by using ASP.NET MVC, you won’t be able to have a rich UI. In truth, you won’t only have access to a wider range of controls; you’ll also avoid a bunch of poor abstractions, and generate JavaScript by writing JavaScript, css by writing css and html by writing html.
4 – Browser differences handled for you
I’m guessing that the claim is that some of the controls mentioned in point 3 might render different HTML based on the requesting browser. Guess what, most jQuery (or any other js framework) plug-ins are fully compatible with all relevant browsers because they too can generate different HTML. In fact, doing this on the client side is almost always better – since you can tell the exact capabilities of a browser.
Also, it would probably be better if you generated correct HTML, CSS and JS in the first place – something you can’t normally control using ASP.NET WebForms. So not only is this really just a benefit because IE is a pain, but its only worth mentioning because points #1, #2 and #3 mean that you’ve lost complete control over doing it right in the first place.
5 – SharePoint builds on WebForms
Yes, if you use SharePoint, you’ll have to use WebForms.
Now on to ASP.NET MVC:
1 – Feels comfortable for many traditional web developers
ASP.NET WebForms is familiar while ASP.NET MVC is comfortable – that’s helpful. Nonetheless, when I see “traditional” I think “not-modern”. A more honest counterpoint to the WebForms claim would be: “A more natural way to build web applications”. WebForms tried to help WinForm developer’s transition to the web. ASP.NET MVC is a model that better reflect the realities of programming on the web. It’s more than just comfort, and has nothing to do with tradition.
2 – Total control of HTML markup
HTML, JS and CSS are yours to command. That doesn’t mean you can’t use controls to speed up development and improve your applications. The way this is worded sure makes it sound like MVC is a lot more work than WebForms though. It isn’t.
3 – Supports Unit Testing, TDD and Agile methodologies
I’m not sure what the technology stack has to do with the development methodology, so we’ll just ignore the last part. That aside, its true that MVC makes it possible to unit test your code. The counter point to that is that WebForms is essentially impossible to unit test. This also understates the architectural superiority and design of MVC – its doesn’t only allow you to leverage a number of best practices, it itself is actually built around those same practices. Code that can be unit tested, regardless of whether it is or not, is almost always superior to code that cannot.
4 – Encourages more prescriptive applications
So if ASP.NET MVC lets you build you application the way you should be building it, should you infer that ASP.NET WebForm forces you to build applications the wrong way? Yes, you should.
5 – Extremely flexible and extensible
Both frameworks share this value – but ASP.NET MVC is more about building on top of existing code, while ASP.NET WebForms is more about hacking things until they work. If you think this means that ASP.NET MVC can only be useful once you’ve extended it, then you are wrong. It works great out of the box is is feature rich.
Other Stuff
The video goes on to make weird assertions, like the possibility of turning back and picking a different stack if you feel you’ve made the wrong choice because of how similar and how much infrastructure they share. The better solution is to pick the right technology because going back months or years into your project doesn’t sound like good advice to me.
It also mentions that it’s common to have some pages handled by MVC and others by WebForms. It’s good to know that you can do this – especially since it’s a good way to upgrade from WebForms to MVC. However, I’d hardly call it common or even recommended. It’s a useful transitional tool which you should aim to get out of as quickly as possible.
Ultimately, the first 4 values of WebForms all boil down to the same thing: there’s a
System.Web.UI namespace which represents the wrong way to build a web app. There are good reasons to pick WebForms – but they all come down to time and practicalities of learning new things (and SharePoint). I won’t tell you that you have to learn MVC, because that may not be practical for you. I’ll repeat what I’ve said before, ASP.NET MVC and WebForms DO NOT serve different purposes and one is not better suited for a particular type of application than the other(except SharePoint). They are completely overlapping technologies, and ASP.NET MVC is superior to WebForms.
MVC was only “formally” released early last year. There are very few 3rd party components made for MVC, compared to ASP.Net. In order to support MVC, you would have to find and hire persons who know MVC. MVC is simply not widely used yet. There is nothing that I cannot be done with ASP.Net. Why would I change something that’s not broken?
You indicated that this article would discuss the pros and cons of each. All you’ve done is try to sell us on MVC and slam ASP.Net. Can you point out one place where you indicated the pros of ASP.Net — where you actually thought that ASP.Net was better? Yah.. I didn’t think so.
You’re so in bed with MVC. Go take a shower man
Dragmonkeys need to be shot.
OK, I’m kidding. Let me rephrase that:
Dragmonkeys need some HTML lessons. And then some MVC books.
I think they’re both sub-optimal. With enough prodding and coaxing you can bend either to your will. In the hands of a qualified developer, I don’t see one being any better than the other. I’ve been serving up HTML since the IDC/HTX days and before that using CGI executables and scripts. It’s no way to build an “application”. Bring on the client side Javas! er, um… I mean Silverlights! HTML5 by 2015? 2025? Any bets?
Comparisons aside, use the tools which pay your bills. Right now all the ASP.NET open positions in my area are WebForms and not MVC. However learn MVC on the side. You will use it sometime in the future when it becomes more ubiquitous.
Chris,
But this isn’t how Web Forms is being sold to the public. The “benefits” of Web Forms that is constantly being peddled by Microsoft (and fans of Web Forms) is the ability to using complex controls and the ability to create a thick client.
I completely agree in the YAGNI world if there is a requirement to build a one page site that simply spits out some data for internal use only, there’s nothing wrong with creating a Web Form app to do so. I did just that a few weeks ago and I loathe Web Forms but I needed to hit a DB and display the contents of a single table in a grid/table. No style sheets, no other controls, just a page where a table dump could be read by the Hardware Developers. For that job Web Forms did the job in minutes, and I was thankful for it.
The problem is, that frequency of projects like that is few and far in between.
While I definitely prefer MVC and advocate it’s use over WebForms on a daily basis, I’d have to agree with Pete. If you need to create a pretty simple application that just does some database interaction (i.e. display a table, allow you to edit the data) you can create it in WebForms much faster than you can in MVC. The output it not going to be something anyone would consider a “Quality web site” but if WebForms gets the job done and fulfills all the needs of the users its still a viable option.
Last 4 monthes I working with ASP.NET MVC in my free time, and with classic ASP.NET in my working time, and find MVC is much more better. It allows to create better user interface with javascript, easy to use, but more difficult to understand.
Very well said again, Karl!
Bertrand raises an important point above, saying that for “web controls … real reason is to provide higher level abstractions than div and input”.
I would agree with that, but still thinks that it is insane to generate client code on a server, trying to predict all the features and quirks of client-side interface and logic.
To me, a proper design includes client side tools/libraries and only *very basic* server side generators which together allow to avoid/diminish amount of low-level (‘div’ and ‘input’) client code manipulations.
Business logic is largely contained on a server.
This makes extremely [most] important to have a standard generic protocol and transport mechanism (think about JSON) allowing to pass information / requirements between server and client.
For example, there should be a set of rules which allow server to tell client what validation it has to perform on user-entered data.
If server and client can speak and understand each other (i.e. common protocol), there is no need any more to generate the whole client data validation code on server. Instead server just tells client what to check for.
“I agree, but your average corporate developer doesn’t want to learn jQuery” or DI, or SOLID, or anything really. That’s why your average corporate developer, the people you are defending, aren’t worth the keystrokes it took to type this.
It’s the embodiment of what is wrong with what we do and helpful little videos like this make it worse.
I think you’re right but your point would have been better made had you “mis de l’eau dans ton vin” (tone it down).
With that being said, good post as usual!
Justin,
I’m not saying it’s not done, but I’d be surprised if it was anywhere close to common.
I just find it weird that a site like would even post such a video, I guess it’s to get hits or something since it’s just flame bait anyways.
See, just look at our Bullshit Detector, he already knows that Karl is out of touch…apparently he migrated here from WekeRoad.com.
I would guess that the argument that Microsoft is trying to make is one that doesn’t involve putting a bullet in the head of WebForms. I think they did that before with VB and that didn’t go that well for them. Doesn’t do any harm to point out better ways of doing WebForms while people are still using it.
Anyway, we all know that ASP.NET MVC sucks too. FubuMVC and OpenRasta FTW!
@Andrew,
I don’t know how common it is but I am running MVC alongside of Webforms. I didn’t really have a choice. I wanted to get out of Webforms but I was not going to rewrite 3 years worth of code. It was actually fairly easily to get MVC and WEbforms working together. Now any new code gets written in MVC and I can choose to rewrite and old webforms code if and when I want to.
@Ben and Onur:
Your argument, re MVP frameworks that sit on top of WebForms is valid. But that isn’t the argument Microsoft is trying to make. And the vast majority of people who are going to pick one or the other are going to be using as-is out of the box.
Hi Karl,
what a load of arrogant, self-righteous crap:
‘should you infer that ASP.NET WebForm forces you to build applications the wrong way? Yes, you should.’
Obviously written by someone who is out of touch with the reality of developing software for a living.
I agree with you Karl. To give you example of why we should learn and adopt MVC is : What happens to webforms when HTML 5 is popular suddenly (with IE 9 around, I have no doubt even Microsoft will move in that direction). All those custom/3rd party webform controls and even some of Microsoft’s own components like GridView will take some time catching up.
It goes beyond saying that even as of today any one that has anything to do with Web Development must know ins and outs of HTML,CSS and javascript.
I am not saying Webforms is bad, but I don’t see any reason why I should start anything new in that. Picking up MVC is extremely simple for experienced Web forms guys. I myself did it in a week
.
“Would you consider generating your HTML from a stored procedure?”
What’s so wrong with this??
“The counter point to that is that WebForms is essentially impossible to unit test. This also understates the architectural superiority and design of MVC ”
Oh yeah ? And apparently you never heard of MVP pattern.
At least in this post you are not cursing as your previous MVC , Webforms post
@Eric
Same here. I’ve been using MVC a lot lately and web forms are pretty bad. I know a case can be made to use them in certain situations, but right now I want nothing to do with em.
I’ve been slowly coming to the same conclusions as you, and every day I spend working with MVC I love it (MVC) more and loathe Web Forms more.
I ranted about my love/hate affair with Web Forms a couple of weeks ago:
Wayne,
More often than not, that sort of thinking leads to no good. It leads to very thick web clients which are buggy as hell, take forever to load (due to about 9 pages of ViewState) and results in a maintenance nightmare.
If your client/customer/management wants a think client, why not just propose one? A website is a website, it can have some client side goodness (basic validation, etc.), but if you find yourself getting requirements like “Make the grid act just like Excel”, then its time to step back and realized that maybe the web should be used to launch a Winforms/WCF/Silverlight application, not do a poor job in implementing a thick client.
Ah, while I’m firmly in the MVC camp myself (works for me is the only argument I need), I think your point number two calls for comments.
The point of web controls is not to hide the HTML away. That has always been fallacious: if that were the intention, how about the rest of the page? Why isn’t the whole of WebForms pages written in a language that’s not HTML?
No, the real reason is to provide higher level abstractions than div and input. Granted, there are more modern approaches now but by claiming that it’s “insane” to think controls generating HTML have any advantage over plain HTML/CSS/JS, I think you are building a straw-man, among other logical fallacies.
@Troy:
I agree, but your average corporate developer doesn’t want to learn jQuery, they really do want to drag+drop some third party UI, make a few changes to the properties via a wizard dialog, and be done with it.
@Karl:
Agreed as well. I’m learning ASP.NET MVC at the moment, but most of the jobs I’ve seen/applied for/talked to others about are still heavily WebForms.
@Wayne
I’m on the fence….I’ve seen that shortsightedness for intranet “thick client” apps bite people many times. I hear the same argument about silverlight, and I’m just not a believer.
@peter
Can you redistribute system.web.mvc.dll now or not? If not I agree its a problem.
As for speed, YMMV, but i know I’m faster now on MVC than I was at my peak on WebForms – but that might be due to other, unrelated growth.
I still think my myopic stances makes for better reading, and results in more meaningful discussion. Its certainly fun to write.
I know django well enough…and I do like it.
How do you create a setup wizard for an MVC web app that installs to IIS 6?
If you have a well-built existing app and you need to build out a web interface as quickly as possible, do you think you can write it faster in MVC than a few web forms and some devxpress controls?
I’m just playing devil’s advocate, but my point is both of these tools should be taken into consideration.
I respect your writing and expertise, but sometimes you take a very extreme myopic stance, such as “dont use try/catch”.
If your standpoint on web apps is based solely on cleanliness, then you should look at django or turbogears, which are far cleaner than MVC.
Wayne Molina:
“make use of a lot of VERY nice third party UI components”
There are a whole lot MORE UI components built in server-agnostic client side libraries like jQuery that will be much easier to use when developing with MVC.
Sure Microsoft has to defend this antiquated paradigm of web programming that many of their products (which are as antiquated as WebForms… cough: SharePoint) use.
I’m actually somewhat surprised that something like ASP.Net MVC could actually come from Microsoft however to be fair there was a lot of collaboration with the community. Actually Karl was being easy on Microsoft.
I still maintain that WebForms are good for developing what would amount to a thick-client app (e.g. a Corporate CRM or Time Tracking system) because you can do it relatively painlessly and make use of a lot of VERY nice third party UI components. To the average corporate developer these are good things.
ASP.NET MVC is better overall because of the flexibility and control that you get, but there are situations where that flexibility doesn’t really matter, and RAD/using those $1K controls the director bought is the concern.
As always, both are tools and one isn’t “better” than the other. I daresay that for the majority of professional .NET developers out there, and the majority of .NET work, WebForms are the preferred solution over MVC.
There goes Karl again, bashing Microsoft.
There, just wanted to get that out of the way because a comment like that will be coming sooner or later.
Can’t say I disagree with anything you said, I guess I’m just surprised that would put this video up on their site. Common to run MVC alongside Web Forms? Sure…
While I agree with your overall conclusion (ASP.NET MVC is overlapping and superior), it is not the case that WebForms apps have to be untestable or (always) about “hacking things until they work”. We are happily testing and running a large webforms app using our own simple MVP framework and the Readify guys in Aus are doing the same with their framework.
IMO – A lot of the crap that was done with WebForms historically was because as a community we were still writing crap procedural code. Pre-SOLID, pre-IOC, pre-ProperOO (POO?)
And yes, I know the WF framework does not make achieving these things easy.
Again, I’m not being a WebForms apologist. I think the WebForms abstraction is an abomination and I would never personally use it again when given the choice. However, it is here to stay as we all know MS are not going to say “Actually, yeah, webforms suck so all your investments are worthless” anytime soon (regardless of privately held opinions). In the meantime we need to help the folks still in WF land (and there are a lot of them stuck there).
I couldn’t agree more with all of this. I have always disliked the arguments that WebForms is something good just because it “kinda sorta” resembles WinForms (which is a lousy WIN32 wrapper and should never, never ever, be used, now that WPF exists). | http://codebetter.com/karlseguin/2010/03/11/webforms-vs-mvc-again/?utm_source=feedburner&utm_medium=feed&utm_campaign=Feed%3A+CodeBetter+%28CodeBetter.Com%29 | crawl-003 | en | refinedweb |
SYNOPSIS
#include <curses.h>
DESCRIPTION (patch 20090803).. Differ-
ences from the SVr4 curses are summarized under the EXTENSIONS and
PORTABILITY sections below and described in detail in the respective
EXTENSIONS,
/etc/terminfo/a/att4424.vline border(3NCURSES)
wvline_set border_set(3NCURSES)NCURSES)). You may
set the foreground and background color values with this environ-
ment variable by proving a 2-element list: foreground,background.
For example, to tell ncurses to not assume anything about the col-
ors,., /etc
Several different configurations are possible, depending on the config-
ure script options used when building ncurses. There are a few main
options whose effects are visible to the applications developer using
ncurses:
/usr/share/tabset
directory containing initialization files for the terminal capa-
bility database /etc/terminfo terminal capability database
SEE ALSO
terminfo(5) and related pages whose names begin "curs_" for detailed
routine descriptions.
EXTENSIONS
ingNCURSES) manual page for details.
The ncurses library includes a function for directing application out-
put to a printer attached to the terminal device. See the
print(3NCURSES) manual page for details.
PORTABILITY
The imple-
mentation. appli-
cation programs. See opaque(3NCURSES) for the discussion of
is_scrollok, etc.
In historic curses versions, delays embedded in the capabilities cr,
This was an undocumented feature of AT&T System V Release 3 curses.
AUTHORS
Zeyd M. Ben-Halim, Eric S. Raymond, Thomas E. Dickey. Based on pcurses
by Pavel Curtis. | http://www.linux-directory.com/man3/ncurses.shtml | crawl-003 | en | refinedweb |
SYNOPSIS
#include <time.h>
int nanosleep(const struct timespec *rqtp, struct timespec *rmtp);
DESCRIPTION
The nanosleep() function shall cause sys-
tem. But, except for the case of being interrupted by a signal, the
suspension time shall not be less than the time specified by rqtp, as
measured by the system clock CLOCK_REALTIME.
The use of the nanosleep() function has no effect on the action or
blockage of any signal.
RETURN VALUE
If the nanosleep() function returns because the requested time has
elapsed, its return value shall be zero.
If the nanosleep() function returns because it has been interrupted by
a signal, it shall return a value of -1 and set shall return a value of -1 and set errno to
indicate the error.
signal number. This volume of.
FUTURE DIRECTIONS
None.
SEE ALSO
sleep() , . | http://www.linux-directory.com/man3/nanosleep.shtml | crawl-003 | en | refinedweb |
JBoss.orgCommunity Documentation
Version: 5.1.0.trunk
Drools is a business rule management system (BRMS) with a forward chaining inference based rules engine, more correctly known as a production rule system, using an enhanced implementation of the Rete algorithm.
In this guide we are going to get you familiar with Drools Eclipse plugin which provides development tools for creating, executing and debugging Drools processes and rules from within Eclipse.
It is assumed that you has some familiarity with rule engines and Drools in particular. If no, we suggest that you look carefully through the Drools Documentation.
Drools Tools come bundled with JBoss Tools set of Eclipse plugins. How to install JBoss Tools you can find in the Getting Started Guide.
The following table lists all valuable features of the Drools Tools.
The latest JBossTools/JBDS documentation builds
All JBoss Tools/JBDS documentation you can find on the documentation release page.
In this chapter we are going to show you how to setup an executable sample Drools project to start using rules immediately.
First, we suggest that you use Drools perspective which is aimed at work with Drools specific resources.
To create a new Drools project follow to File > New > Drools Project. This will open New Drools Project wizard like on the figure below.
On the first page type the project name and click Next.
Next you have a choice to add some default artifacts to it like sample rules, decision tables or ruleflows and Java classes for them. Let's select first two check boxes and press Next.
Next page asks you to specify a Drools runtime. If you have not yet set it up, you should do this now by clicking the Configure Workspace Settings link.
You should see the Preferences window where you can configure the workspace settings for Drools runtimes. To create a new runtime, press the Add button. The appeared dialog prompts you to enter a name for a new runtime and a path to the Drools runtime on your file system.
A Drools runtime is a collection of jars on your file system that represent one specific release of the Drools project jars. While creating a new runtime, you must either point to the release of your choice, or you can simply create a new runtime on your file system from the jars included in the Drools Eclipse plugin.
Let's simply create a new Drools 5 runtime from the jars embedded in the Drools Eclipse plugin. Thus, you should press Create a new Drools 5 runtime button and select the folder where you want this runtime to be created and hit OK.
You will see the newly created runtime show up in your list of Drools runtimes. Check it and press OK.
Now press Finish to complete the project creation.
This will setup a basic structure, classpath and sample rules and test case to get you started.
Now let's look at the structure of the organized project. In the Package Explorer you should see the following:
The newly created project contains an example rule file Sample.drl in the src/main/rules directory and an example java file DroolsTest.java that can be used to execute the rules in a Drools engine in the folder src/main/java , in the com.sample package. All the others jar's.
Now we are going to add a new Rule resource to the project.
You can either create an empty text .drl file or make use of the special New Rule Resource wizard to do it.
To open the wizard follow to File > New > Rule Resource or use the menu with the JBoss Drools icon on the toolbar.
On the wizard page first select /rules as a top level directory to store your rules and type the rule name. Next it's mandatory to specify the rule package name. It defines a namespace that groups rules together.
As a result the wizard generates a rule skeleton to get you started.
This chapter describes how to debug rules during the execution of your Drools application.
At first, we'll focus on how to add breakpoints in the consequences of your rules.
Whenever such a breakpoint is uncounted during the execution of the rules, the execution is halted. It's possible then inspect the variables known at that point and use any of the default debugging actions to decide what should happen next (step over, continue, etc). To inspect the content of the working memory and agenda the Debug views can be used.
You can add/remove rule breakpoints in .drl files in two ways, similar to adding breakpoints to Java files:
Double-click the ruler in the Rule.
Right-click the ruler. Select Toggle Breakpoint action in the appeared popup menu. Clicking the action will add a breakpoint at the selected line or remove it if there is one already.
The Debug perspective contains a Breakpoints view which can be used to see all defined breakpoints, get their properties, enable/disable or remove them, etc. You can switch to it by navigating to Window > Perspective > Others > Debug.
Drools breakpoints are only enabled if you debug your application as a Drools Application. To do this you should perform one of the actions:
Select the main class of your application. Right click it and select Debug As > Drools Application.
Alternatively, you can also go to Debug As > Debug Configuration to open a new dialog for creating, managing and running debug configurations.
Select the Drools Application item in the left tree and click the New launch configuration button (leftmost icon in the toolbar above the tree). This will create a new configuration and already fill in some of the properties (like the Project and Main class) based on main class you selected in the beginning. All properties shown here are the same as any standard Java program.
Remember to change the name of your debug configuration to something meaningful.
Next click the Debug button on the bottom to start debugging your application.
After enabling the debugging, views can also be used to determine the contents of the working memory and agenda at that time as well (you don't have to select a working memory now, the current executing working memory is automatically shown).
A domain-specific language is a set of custom rules, that is created specifically to solve problems in a particular domain and is not intended to be able to solve problems outside it. A DSL's configuration is stored in plain text.
In Drools this configuration is presented by .dsl files that can be created by right click on the project->New->Other->Drools->Domain Specific Language.
DSL Editor is a default editor for .dsl files:
In the table below all the components of the DSL Editor page are described:
This wizard can be opened by double clicking some line in the table of language message mappings or by clicking the Editbutton.
On the picture below you can see all the options,Edit language mapping Wizard allow to change.
Their names as well as the meaning of the options are correspond to the rows of the table.
To change the mapping a user should edit the otions he want and finally click Ok.
This wizard is equal to Edit language mapping Wizard. It can be opened by clicking the Add button.
The only difference is that instead of editing the information you should enter new one.
Drools tools also provide some functionality to define the order in which rules should be executed.Ruleflow file allows you to specify the order in which rule sets should be evaluated using a flow chart. So you can define which rule sets should be evaluated in sequence or in parallel as well as specify conditions under which rule sets should be evaluated.
Ruleflows can be set only by using the graphical flow editor which is part of the Drools plugin for Eclipse. Once you have set up a Drools project,you can start adding ruleflows. Add a ruleflow file(.rf) by clicking on the project and selecting "New -> Other...->Flow File":
By default these ruleflow files (.rf) are opened in the graphical Flow editor. You can see it on the picture below.
The Flow editor consists of a palette, a canvas and an outline view. To add new elements to the canvas, select the element you would like to create in the palette and then add it to the canvas by clicking on the preferred location.
Clicking on the Select option in the palette and then on the element in your ruleflow allows you to view and set the properties of that element in the properies view.
Outline View is useful for big complex schemata where not all nodes are seen at one time. So using your Outline view you can easly navigate between parts of a schema.
Flow editor supports three types of control elements. They are:
The Rule editor works on files that have a .drl (or .rule in the case of spreading rules across multiple rule files) extension.
The editor follows the pattern of a normal text editor in eclipse, with all the normal features of a text editor:
While working in the Rule editor you can get a content assistance the usual way by pressing Ctrl + Space.
Content Assist shows all possible keywords for the current cursor position.
Content Assist inside of the Message suggests all available fields.
Code folding is also available in the Rule editor. To hide/show sections of the file use the icons with minus/plus on the left vertical line of the editor.
The Rule editor works in synchronization with the Outline view which shows the structure of the rules, imports in the file and also globals and functions if the file has them.
The view is updated on save. It provides a quick way of navigating around rules by names in a file which may have hundreds of rules. The items are sorted alphabetically by default.
The Rete Tree view shows you the current Rete Network for your .drl file. Just click on the Rete Tree tab at the bottom of the Rule editor.
Afterwards you can generate the current Rete Network visualization. You can push and pull the nodes to arrange your optimal network overview.
If you got hundreds of nodes, select some of them with a frame. Then you can pull groups of them.
You can zoom in and out the Rete tree in case not all nodes are shown in the current view. For this use the combo box or "+" and "-" icons on the toolbar.
The Rete Tree view works only in Drools Rule Projects, where the Drools Builder is set in the project properties.
We hope, this guide helped you to get started with the JBoss BPMN Convert module. Besides, for additional information you are welcome on JBoss forum. | http://docs.jboss.org/tools/3.1.0.CR2/en/drools_tools_ref_guide/html_single/index.html | crawl-003 | en | refinedweb |
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cant see namespaces in my code
prilla
July 2nd, 2002, 08:25 AM
Hi,
I am learning asp.net and I have visual studio enterprise architect installed. As a prerequisite I had teh .net framework installed. the os is win2k server. it wont install active directory but this is because (i think) i dont have a network card.
the problem is this:
when I run a sample application in visual studio.net, that sends an email when u enter your name and email address on a form and click send button, it comes up with an error - cant override the new SMTPMail () method as it is private??? This is a sample application so it should run fine. the system.web.mail namespace is imported and this contains the smtpmail () method which is a public method, not private. i had the same problem running another sample application that uses ado.net. it seems like the class library is the .net framework cannot be seen by visual studio .net...
do i need to specify its location as a classpath environment variable like you would with java? or is there something i missed...
thanks!
Priscilla
codeguru.com | http://forums.codeguru.com/archive/index.php/t-197279.html | crawl-003 | en | refinedweb |
for connected embedded systems
SyncMutexLock(), SyncMutexLock_r()
Lock a mutex synchronization object
Synopsis:
#include <sys/neutrino.h> int SyncMutexLock( sync_t * sync ); int SyncMutexLock_r( sync_t * sync );
Arguments:
- sync
- A pointer to the synchronization object for the mutex that you want to lock.
Library:
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
Description:
- STATE_MUTEX
- The calling thread blocks waiting for the synchronization object to be unlocked.
Returns:
The only difference between these functions is the way they indicate errors:
- SyncMutexLock()
- If an error occurs, the function returns -1 and sets errno. Any other value returned indicates success.
- SyncMutexLock_r()
- Returns EOK on success. This function does NOT set errno. If an error occurs, this function returns any value listed in the Errors section.
Errors:
- EAGAIN
- On the first use of a statically initialized sync, all kernel synchronization objects were in use.
- EFAULT
- A fault occurred when the kernel tried to access the buffers you provided.
- EINVAL
- The synchronization ID specified in sync doesn't exist.
- ETIMEDOUT
- A kernel timeout unblocked the call. See TimerTimeout().
Classification:
See also:
pthread_mutex_lock(), pthread_mutex_unlock(), SyncTypeCreate(), SyncDestroy(), SyncMutexUnlock() | http://www.qnx.com/developers/docs/6.4.0/neutrino/lib_ref/s/syncmutexlock.html | crawl-003 | en | refinedweb |
for connected embedded systems
SyncDestroy(), SyncDestroy_r()
Destroy a synchronization object
Synopsis:
#include <sys/neutrino.h> int SyncDestroy( sync_t* sync ); int SyncDestroy_r ( sync_t* sync );
Arguments:
- sync
- The synchronization object that you want to destroy.
Library:
libc
Use the -l c option to qcc to link against this library. This library is usually included automatically.
Description:.
Blocking states
These calls don't block.
Returns:
The only difference between these functions is the way they indicate errors:
- SyncDestroy()
- If an error occurs, the function returns -1 and sets errno. Any other value returned indicates success.
- SyncDestroy_r()
- Returns EOK on success. This function does NOT set errno. If an error occurs, the function can return any value listed in the Errors section.
Errors:
- EBUSY
- The synchronization object is locked by a thread.
- EFAULT
- A fault occurred when the kernel tried to access sync.
- EINVAL
- The synchronization ID specified in sync doesn't exist.
Classification:
See also:
pthread_cond_destroy(), pthread_mutex_destroy(), pthread_rwlock_destroy(), sem_destroy(), SyncTypeCreate() | http://www.qnx.com/developers/docs/6.4.0/neutrino/lib_ref/s/syncdestroy.html | crawl-003 | en | refinedweb |
Operating systems, development tools, and professional services
for connected embedded systems
for connected embedded systems
SyncCtl(), SyncCtl_r()
Perform an operation on a synchronization object
Synopsis:
#include <sys/neutrino.h> int SyncCtl( int cmd, sync_t * sync, void * data ); int SyncCtl_r( int cmd, sync_t * sync, void * data );
or
- attach an event to a mutex so you'll be notified when the mutex changes to the DEAD state.
These functions are similar, except for the way they indicate errors. See the Returns section for details.
Returns:
The only difference between these functions is the way they indicate errors:
- SyncCtl()
- If an error occurs, the function returns -1 and sets errno. Any other value returned indicates success.
- SyncCtl or data.
- EINVAL
- The synchronization object pointed to by sync doesn't exist, or the ceiling priority value pointed to by data is out of range.
- ENOSYS
- The SyncCtl() and SyncCtl_r() functions aren't currently supported.
Classification:
See also:
pthread_mutex_getprioceiling(), pthread_mutex_setprioceiling(), SyncCondvarSignal(), SyncCondvarWait(), SyncDestroy(), SyncMutexEvent(), SyncMutexLock(), SyncMutexRevive(), SyncMutexUnlock(), SyncTypeCreate() | http://www.qnx.com/developers/docs/6.4.0/neutrino/lib_ref/s/syncctl.html | crawl-003 | en | refinedweb |
fwprintf, swprintf, wprintf - print formatted wide-character output
Synopsis
Description
Return Value
Errors
Examples
Application Usage
Rationale
Future Directions
See Also
#include <stdio.h>
#include <wchar.h>
int fwprintf(FILE *restrict stream, const wchar_t *restrict format, ...);
int swprintf(wchar_t *restrict ws, size_t n,
const wchar_t *restrict format, ...);
int wprintf(const wchar_t *restrict format, ...);
The fwprintf() function shall place output on the named output stream. The wprintf() function shall place output on the standard output stream stdout. The swprintf() function shall place output followed by the null wide character in consecutive wide characters starting at *ws; no more than n wide characters shall be written, including a terminating null wide character, which is always added (unless n is zero).
Each of these functions shall convert, format, and print its arguments under control of the format wide-character string. The format is composed of zero or more directives: ordinary wide-characters, which are simply copied to the output stream, and conversion specifications, each of which results in the fetching of zero or more arguments. The results}], giving the position of the argument in the argument list. This feature provides for the definition of format wide-character strings that select arguments in an order appropriate to specific languages (see the EXAMPLES section).
The format can contain either numbered argument specifications (that is, "%n$" and "*m$"), or unnumbered argument conversion specifications (that is, % and * ), but not both. The only exception to this is that %% can be mixed with the "%n$" form. The results of mixing numbered and unnumbered argument specifications in a format wide-character string are undefined. When numbered argument specifications are used, specifying the Nth argument requires that all the leading arguments, from the first to the (N-1)th, are specified in the format wide-character string.
In format wide-character strings containing the "%n$" form of conversion specification, numbered arguments in the argument list can be referenced from the format wide-character string as many times as required.
In format wide-character strings containing the % form of conversion specification, each argument in the argument list shall be used exactly once.
All forms of the fwprintf() function allow for the insertion of a locale-dependent radix character in the output string, output as a wide-character value. The radix character is defined in the programs locale (category LC_NUMERIC ). In the POSIX locale, or in a locale where the radix character is not defined, the radix character shall default to a period ( . ).
Each conversion specification is introduced by the % wide character or by the wide wide-character:
wprintf(L"%1$d:%2$.*3$d:%4$.*3$d\n", hour, min, precision, sec);
The flag wide characters and their meanings are:If a conversion specification does not match one of the above forms, the behavior is undefined.
In no case does a nonexistent or small field width cause truncation of a field; if the result of a conversion is wider than the field width, the field shall be expanded to contain the conversion result. Characters generated by fwprintf() and wprintf() shall be printed as if fputwc() had been called.wprintf() or wprintf() and the next successful completion of a call to fflush() or fclose() on the same stream, or a call to exit() or abort().
Upon successful completion, these functions shall return the number of wide characters transmitted, excluding the terminating null wide character in the case of swprintf(), or a negative value if an output error was encountered, and set errno to indicate the error.
If n or more wide characters were requested to be written, swprintf() shall return a negative value, and set errno to indicate the error.
For the conditions under which fwprintf() and wprintf() fail and may fail, refer to fputwc() .
In addition, all forms of fwprintf() may fail if:The following sections are informative.
To print the language-independent date and time format, the following statement could be used:
wprintf(format, weekday, month, day, hour, min);
For American usage, format could be a pointer to the wide-character string:
L"%s, %s %d, %d:%.2d\n"
producing the message:
Sunday, July 3, 10:02
whereas for German usage, format could be a pointer to the wide-character string:
L"%1$s, %3$d. %2$s, %4$d:%5$.2d\n"
producing the message:
Sonntag, 3. Juli, 10:02
None.
None.
None.
btowc() , fputwc() , fwscanf() , mbrtowc() , setlocale() , . | http://www.squarebox.co.uk/cgi-squarebox/manServer/usr/share/man/man3p/swprintf.3p | crawl-003 | en | refinedweb |
Games, .NET, Performance, and More!
The program I used to replicate this behavior is shown below. I had to put the WriteLine's in, because the /optimize+ versions were omitting the n = i statement, since n was never used. While I found that portion of the code extremely smart, I found that the variable re-usage of all forms of compilation were extremely lacking. For example, the IL output from each function is as follows:ShowBlockScopedVariablesInILAsMethodScoped, this method demonstrates that loop scoped variables within a C# application, actually manifest at the method level rater than somehow manifesting at the loop scoped level. This makes perfect sense so that upon entry into the method, the stack is set to the appropriate size as needed by the running code. 2 stack variables are allocated in this method as expected.
ShowMultipleBlockScopedVariablesInILAsMultipleMethodScopedVariables, this method demonstrates that loop scoped variables within a C# application don't get re-used at all. Normally, you would expect any stack variable used in a loop, to become available for use by the program in another loop once the current loop has been exited. However, this isn't the case since there are 6 stack variables allocated 2 for each loop with one variable mapping to i in each loop and the other being n.
ShowMethodScopedVariablesReUsed, this method demonstrates that you can reduce your stack footprint by not using loop scoped variables and simply declaring method scoped variables and re-using them manually. This is somewhat counter-intuitive to me, since the language could optimize the loop scoped variables and re-use them, but doesn't.
Long story short, try not to use locally scoped variables for loop constructs unless you only have one loop in your code. Try to re-use variables where appropriate. This is the lesson I've learned. Maybe I'm trying to be too protective of the generated IL and the JIT is doing something behind the scenes. I notice that the Partition II document shipped with the Framework SDK demonstrates that ILAsm allows locally scoped variables within nested blocks to share the same location as a method scoped variable. You can read about this in 14.4.1.3. I even tried to create some IL that did this and compile it out:
.locals init ( int32 V_0, int32 V_1, [0] int32 V_2, [1] int32 V_3, [0] int32 V_4, [1] int32 V_5)The above compiled just fine under ILAsm. However, it did wind up producing IL that simply removed the remaining 4 variables and made sure the references pointed to V_0 through V_1. However, the really funny thing is that ldloc.{digit} and stloc.{digit} are used everywhere they would have normally been used, but the final two references that mapped V_4/V_5 to V_0/V_1 were listed as stloc.s/ldloc.s. I'm not sure there is any speed increase over any of the different versions of stloc/ldloc, but if there was, even the ILAsm compiler is lacking in performance optimizations.
Well, I'm done with this topic for the evening. I probably won't change very much about how I code based on any of this information. It may force me to use method level variables in places where I have lots of loop structures, but probably not.
using System;
public class LocalityOfVariables { private static void Main(string[] args) { } private static void ShowBlockScopedVariablesInILAsMethodScoped() { for (int i = 0; i < 10; ++i) { int n = i; Console.WriteLine(n); } } private static void ShowMultipleBlockScopedVariablesInILAsMultipleMethodScopedVariables() { for (int i = 0; i < 10; ++i) { int n = i; Console.WriteLine(n); } for (int i = 0; i < 10; ++i) { int n = i; Console.WriteLine(n); } for (int i = 0; i < 10; ++i) { int n = i; Console.WriteLine(n); } } private static void ShowMethodScopedVariablesReUsed() { int i, n; for (i = 0; i < 10; ++i) { n = i; Console.WriteLine(n); } for (i = 0; i < 10; ++i) { n = i; Console.WriteLine(n); } for (i = 0; i < 10; ++i) { n = i; Console.WriteLine(n); } }} | http://weblogs.asp.net/justin_rogers/archive/2004/02/16/73627.aspx#134535 | crawl-003 | en | refinedweb |
#include "OsiSolverInterface.hpp"
#include "CbcModel.hpp"
Include dependency graph for CbcParam.hpp:
Go to the source code of this file.
Parameter codes.
Parameter type ranges are allocated as follows
`Actions' do not necessarily invoke an immediate action; it's just that they don't fit neatly into the parameters array.
This coding scheme is in flux. NODESTRATEGY, BRANCHSTRATEGY, ADDCUTSSTRATEGY, CLEARCUTS, OSLSTUFF, CBCSTUFF are not used at present (03.10.24).
Definition at line 29 of file CbcParam.hpp. | http://www.coin-or.org/Doxygen/CoinAll/_cbc_param_8hpp.html | crawl-003 | en | refinedweb |
On Thu, Sep 19, 2002 at 09:23:46PM +1000, Angus Lees wrote:
> since you're importing, the code/functions will be run in the same
> "namespace" as the page doing the Execute.
>
> thus you can only do $req=shift in either the Executing page or the
> Executed page, since they're shifting the same @_ array.
>
> ie: continue to do $req=shift in your Executing page. remove
> $req=shift from your Executed page, and just continue using $req
> anyway.
I tried that, but it does not work... The values in $req are
undefined as soon as I try to access them in a sub.
This is exactly what I wrote, $req->{log} is defined in the
application object:
[- $req = shift; -]
[- Execute ({inputfile => 'functions.epo', import => 1}); -]
[-
logprint ("bla bla bla"); # <- this is the function
print OUT "my log: $req->{log}"; # here it works
-]
functions.epo:
[$ sub logprint $]
<p>passed values: [+$_[0]+]</p>
<p>logfile: [-print OUT "my log(sub): $req->{log}"; # here not -]</p>
[$ endsub $]
Andre
--
"The inside of a computer is as dumb as hell, but it goes like mad!"
(Physicist Richard Feynman)
---------------------------------------------------------------------
To unsubscribe, e-mail: embperl-unsubscribe@perl.apache.org
For additional commands, e-mail: embperl-help@perl.apache.org | http://mail-archives.apache.org/mod_mbox/perl-embperl/200209.mbox/%3c20020919131512.GE1847@twintower%3e | crawl-003 | en | refinedweb |
ors
SYNOPSIS
#include <curses.h>
int use_default_colors(void);
int assume_default_colors(int fg, int bg);
DESCRIPTION
The use_default_colors() and assume_default_colors() functions fore-
ground:
use_default_colors();
assume_default_colors(-1,-1);
These are ncurses extensions. For other curses implementations, color
number -1 does not mean anything, just as for ncurses before a success-
ful.
RETURN VALUE
These functions return the integer ERR upon failure and OK on success.
They will fail if either the terminal does not support the orig_pair or
orig_colors capability. If the initialize_pair capability is found,
for this application would give unsatisfactory results for a variety of
reasons. This extension was devised after noting that color xterm (and
similar programs) provides a background color which does not necessar-
ily.
PORTABILITY
These routines are specific to ncurses. They were not supported on
Version 7, BSD or System). | http://www.linux-directory.com/man3/assume_default_colors.shtml | crawl-003 | en | refinedweb |
SYNOPSIS
#define _GNU_SOURCE
#include <assert.h>
void assert_perror(int errnum);
DESCRIPTION
If.
abort(3), assert(3), exit(3), strerror(3), feature_test_macros(7)
COLOPHON
This page is part of release 3.23 of the Linux man-pages project. A
description of the project, and information about reporting bugs, can
be found at. | http://www.linux-directory.com/man3/assert_perror.shtml | crawl-003 | en | refinedweb |
#include "GAMSlinksConfig.h"
#include "smag.h"
Include dependency graph for SmagExtra.h:
Go to the source code of this file.
Gives the structure of the Hessian of the objective and each constraint separately.
The user provides space to store the row/col indices and values of hessian entries. The Hessians of all constraints and the objective are stored in one contiguous array. Only the values from the lower-left part of the Hessian are returned. rowStart indicates where the entries for which constraint of the Hessian start. The entries for the objective are found at index rowstart[smagRowCount()]. That is, hesRowIdx[rowStart[connr]..rowStart[connr]-1] are the row indices of the elements in the Hessian of constraint connr if connr<smagRowCount() or the objective if connr==smagRowCount(). Similar for hesColIdx. The Hessian values are computed in the initial level values of the variables and are stored in hesValue. Giving NULL for hesValue switches off the computation of hessian entry values. hesSize should be large enough to store the indices and columns of all Hessians. GAMS uses the estimate 10*prob->gms.nnz*prob->gms.workfactor for the size of the Hessian of the Lagrangian.
EXPERIMENTAL FUNCTION: Computes the directional derivative of the objective function.
Given a vector x and a direction d, computes the function value f(x) and the product <grad f(x), d>. | http://www.coin-or.org/Doxygen/GAMSlinks/_smag_extra_8h.html | crawl-003 | en | refinedweb |
#include </home/clem/local/src/opie/examples/networksettings/examplemodule.h>
Inheritance diagram for VirtualModule:
Definition at line 12 of file examplemodule.cpp.
References m_interfaces, Interface::setHardwareName(), and Interface::setInterfaceName().
Definition at line 25 of file examplemodule.cpp.
References m_interfaces.
[virtual]
Attempts to create a new interface from name you gave possibleNewInterfaces()
Implements Module.
Definition at line 58 of file examplemodule.cpp.
Create and return the Configure Module
Reimplemented from Module.
Definition at line 39 of file examplemodule.cpp.
References l.
Get all active (up or down) interfaces managed by this module. At the end of initialisation you will be asked to return your interfaces Return all of your interfaces even the ones you claimed by isOnwer. Here you can also return your 'virtual' Interface Objects
Definition at line 48 of file examplemodule.cpp.
[inline, virtual]
get the icon name for this device.
Definition at line 24 of file examplemodule.h.
References QString::fromLatin1().
Create, and return the Information Module.
An default Implementation is InterfaceInformationImp
Definition at line 44 of file examplemodule.cpp.
Check to see if the interface i is owned by this module. See if you can handle it. And if you can claim ownership by returning true. For physical devices you will be asked if you want to own the device. But you can also create new
Definition at line 34 of file examplemodule.cpp.
Adds possible new interfaces to the list (Example: usb(ppp), ir(ppp), modem ppp) Both strings need to be translated. The first string is a Shortcut like PPP and the second argument is a description.
list.insert(
list.insert(
Definition at line 52 of file examplemodule.cpp.
References QMap< Key, T >::insert(), and tr.
get dcop calls
Definition at line 25 of file examplemodule.h.
Attempts to remove the interface, doesn't delete i
Definition at line 72 of file examplemodule.cpp.
The current profile has been changed and the module should do any neccesary changes also. As of Opie1.0 profiles are disabled.
Definition at line 16 of file examplemodule.h.
The type of the plugin and the name of the qcop call
Definition at line 15 of file examplemodule.h.
[signal]
Emit this Signal once you change the Interface you're operating on
Reimplemented from Module.
[private]
Definition at line 27 of file examplemodule.h.
Referenced by getInterfaces(), isOwner(), VirtualModule(), and ~VirtualModule(). | http://people.via.ecp.fr/~clem/nist/doxydoc/allOpie/classVirtualModule.html | crawl-003 | en | refinedweb |
#include <EpetraExt_MultiMpiComm.h>
Inheritance diagram for EpetraExt::MultiMpiComm:
Definition at line 54 of file EpetraExt_MultiMpiComm.h.
MultiMpiComm constuctor.
Creates a MultiMpiComm object and communicators for the global and sub- problems.
Definition at line 35 of file EpetraExt_MultiMpiComm.cpp.
MultiMpiComm constuctor, no parallelism over domains.
Creates a MultiMpiComm object for the simple case of no parallelism over multiple steps.
Definition at line 72 of file EpetraExt_MultiMpiComm.cpp.
Copy constructor.
Definition at line 87 of file EpetraExt_MultiMpiComm.cpp.
Destructor.
Definition at line 98 103 of file EpetraExt_MultiMpiComm.cpp.
Definition at line 100 of file EpetraExt_MultiMpiComm.h. | http://trilinos.sandia.gov/packages/docs/r10.0/packages/epetraext/doc/html/classEpetraExt_1_1MultiMpiComm.html | crawl-003 | en | refinedweb |
XGetWindowProperty(3X11) XLIB FUNCTIONS XGetWindowProperty(3X11)
XGetWindowProperty, XListProperties, XChangeProperty, XRota- teWindowProperties, XDeleteProperty - obtain and change win- dow properties
int XGetWindowProperty(Display *display, Window w, Atom pro- perty, for- mat, int mode, unsigned char *data, int nelements); int XRotateWindowProperties(Display *display, Window w, Atom properties[], int num_prop, int npositions); int XDeleteProperty(Display *display, Window w, Atom pro- perty);
actual. Pos- sible values are 8, 16, and 32. This information allows the X server to correctly perform byte-swap operations as necessary. If the format is 16-bit or 32-bit, you must explicitly cast your data pointer to an (unsigned char *) in the call to XChangeProperty. XFree86 Version 4.5.0 1 XGetWindowProperty(3X11) XLIB FUNCTIONS XGetWindowProperty(3X11)Ap- pend.Win- dowProperty. w Specifies the window whose property you want to obtain, change, rotate or delete.
The XGetWindowProperty function returns the actual type of the property; the actual format of the property; the number of 8-bit, 16-bit, or 32-bit items transferred; the number of XFree86 Version 4.5.0 2 XGetWindowProperty(3X11) XLIB FUNCTIONS XGetWindowProperty(3X11) bytes remaining to be read in the property; and a pointer to the data actually returned. XGetWindowProperty sets the return arguments as follows: + If the specified property does not exist for the speci- fied argu- ment. The nitems_return argument is empty. + If the specified property exists and either you assign AnyPropertyType to the req_type argument or the speci- fied type matches the actual property type, XGetWin- dow pro- perty (indexing from zero), and its length in bytes is L. If the value for long_offset causes L to be nega- tive, ele- ments. XFree86 Version 4.5.0 3 XGetWindowProperty(3X11) XLIB FUNCTIONS XGetWindowProperty(3X11) XGetWindowProperty always allocates one extra byte in prop_return (even if the property is zero length) and sets it to zero so that simple properties consisting of charac- ters do not have to be copied into yet another string before use. If delete is True and bytes_after_return is zero, XGetWin- dowProperty deletes the property from the window and gen- erates Pro- perty. XFree86 Version 4.5.0 4 XGetWindowProperty(3X11) XLIB FUNCTIONS XGetWindowProperty(3X11) proper- tiesi- tions). If npositions mod N is nonzero, the X server gen- erates a PropertyNotify event for each property in the order that they are listed in the array. If an atom occurs more than once in the list or no property with that name is defined for the window, a BadMatch error results. If a BadA- tomFree(3X11), XInternAtom. | http://mirbsd.mirsolutions.de/htman/sparc/man3/XDeleteProperty.htm | crawl-003 | en | refinedweb |
Forming Opinions
April 20, 2005
"Tao has reality and evidence but no action or physical form." — Chuang Tzu
Nobody looks forward to forms. Stop a person on the street and ask them what they think about forms and you'll get an earful. Curiously, though, in XML circles forms hold a great deal of interest. Admittedly, not the filling of forms per se, but the technology involved.
Forms and XML have a special affinity. Both are embodiments of structured data exchange. Completing, say, a vacation request by writing on a blank sheet (or Word document) is far less efficient than filling out a form that asks the right questions. Getting rid of paper also proves popular with technology fans. The same holds for data exchange. The right rules and constraints, as embodied in XML syntax and specific vocabulary schemas, yield a vast improvement over an "anything goes" policy of interchange.
Recently, the W3C published a new Member Submission: Web Forms 2.0, or WF2, based on a numbering system where the 1.0 version is the forms chapter of HTML 4.01 plus some DOM interfaces, which I collectively call "classic forms". To be clear, the Submission process is designed to "to propose technology or other ideas for consideration by the Team" — that is, W3C staffers. Unlike documents on the Recommendation track, Submission status doesn't imply any future course for the W3C or any endorsement of the content. It hasn't run the gauntlet of broad participation, review cycles for conformity with the W3C vision, accessibility, international-friendliness, web architecture integration, or Intellectual Property Rights claims — all the things that make official W3C specifications take so long. A Member Submission is just an idea in writing.
WF2 isn't the first forms-related Submission to the W3C. XFDL from UWI (now PureEdge), XFA from JetForm (now Adobe), Form-based Device Input and Upload in HTML from Cisco, and User Agent Authentication Forms from Microsoft were also submitted back in the heady days of 1998 and 1999. These documents were not taken directly to the Recommendation track, vendor wishes notwithstanding, but did provide useful information for what eventually developed as a standard. A segment of the HTML working group considered the existing forms landscape and with W3C approval eventually developed the XForms 1.0 Recommendation.
Exploring Forms
To really get a feel for this new specification, I need to get my hands on some running code as a reference point. I'll start with some basic code and see how WF2 could help. The code is a simple client-side forms framework that I'm calling FormAttic. For full details beyond the brief summary here, I've set up a Wiki page, and am releasing the code under a Creative Commons license.
The client I wrote this for is non-technical, and needs to be able to easily change things around. Thus the key feature of FormAttic is that it uses a declarative technique to record author intent, and is highly amenable to cut-and-paste modification. While it is based on an "executable definition" in JavaScript code, markup could easily be substituted in its place.
In FormAttic, no special markup other than
id attributes is needed on form
controls.
Here
require and
require_one_of are actually functions passed in
to the configuration method
configure_required, which establishes a context for
the declarations. The first parameter of each is a message to be shown if the validation
fails, and subsequent parameters name the controls that are getting marked as required.
Notice that
require_one_of is an instance of a general case of selecting one or
more items from a list — something that requires extra scripting in classic HTML
forms. This format is easily extended to cover additional form declarations, such
as data
types. The remainder of the library consists of initialization code, event handlers
that get
attached in the right places, and auxiliary functions.
The first part of the WF2 document describes the goals and scope of the specification and relationships to existing materials at length. These will be covered in a later installment. For now, let's skip section 1 and get right into the next section — and the code.
Extending Forms
Section 2 starts off saying "At the heart of any form lies the form controls", which seems like a poor choice of words. Many behind-the-scenes structured data exchanges, including tasks currently done with XMLHTTP, can be implemented as a form without controls. The W3C lists current requirement for just this feature. In practice, the primary definition of a form is a model, often as part of a Model-View-Controller pattern; form controls are secondary. It's hard to tell whether this sentence is a minor oversight barely worth mentioning, or something hinting at deeper levels of underlying assumptions. Keep an eye on this as we proceed.
One important class of changes in WF2 consists of newly added attributes that older
clients
will rightly ignore as unknown. Included in this list is the
required
attribute. For FormAttic, this could be implemented by a tiny adjustment to each required
form control:
Browsers that supported WF2 would "just work" with this change. But to support older browsers, an attached script would need to locate all such attributes and configure things such that the form won't submit until the required controls are satisfactorily filled. Essentially, this is what FormAttic script already does. The only difference is where the description of the required state is kept. FormAttic keeps it all in one place, and WF2 spreads it out across the form control markup. Which is better?
The answer, of course, is "it depends". If the document markup is simple or fairly clean, there wouldn't be much difference between the two. On the other hand, in a document riddled with nested tables, font tags, and all manner of non-semantic tag soup, having everything in one place — generally towards the top of the document — is a big plus. Indeed, this was a major motivation for the design of FormAttic.
Another difference is in the DOM interface. Under WF2 the
required state of a
form control would be available for inspection or modification under a
validity
property of the form control object. But again, for older browsers, the script would
need to
do feature detection and in cases where the property isn't implemented, go off and
do its
own thing. Less adroit scripters might even fall back into old habits of browser version
sniffing. On the other hand, for browsers that lack a
validity property, it
would be possible with some care to implement one in client-side script.
Forming an Opinion
A solid case could be made for either all-in-one-place configuration, as in FormAttic, or spread-out-among-form-controls-configuration, as in WF2. The relative benefit of either option is outweighed by the benefit of working with whichever one has wider support in the overall community. My experience pushes me in the direction of keeping things separate, but I'm keeping an open mind and am willing to be persuaded.
As things worked out, I wasn't involved in the production of WF2 to date. As I prepare these columns, I really am forming initial opinions of the technology. Next week, as I continue reviewing WF2, I will focus on the many things I like in the specification.
Births, Deaths, and Marriages
Topologi Difference Detective
A US$29 utility for a broad range of XML-diff actions.
Sun Java Streaming XML Parser
An implementation of JSR, available at java.net.
A first draft of the XML Pipeline specification from Orbeon.
Relax NG seminar in Leuven, Belgium
Relax NG session featuring Eric van der Vlist.
XML Enhancements for Java 1.0
A compiler and runtime system to extend Java 1.4 with first-class support for XML.
Documents and Data
If you walk behind the wall, you see little trees growing behind some of the bricks.
Serious QName advice from Rick Jelliffe.
There should be some kind of award for finding prior-art this fast.
This week's final word on namespaces. | http://www.xml.com/pub/a/2005/04/20/deviant.html | CC-MAIN-2017-13 | en | refinedweb |
Opened 10 years ago
Closed 9 years ago
#4126 closed (worksforme)
TypeError when edit_inline and Multiple ForeignKeys to same model breaks the admin
Description
This is almost a duplicate of #1939, however, I am getting a TypeError instead of a KeyError.
Minimal example (a commenter to #2522 had the same use case in mind):
from django.db import models class Node(models.Model): name = models.CharField('Name', maxlength=30, unique=True) def __str__(self): return self.name class Admin: pass class Link(models.Model): from_node=models.ForeignKey(Node, related_name='has_links_to', edit_inline=True) to_node=models.ForeignKey(Node, related_name='has_links_from', core=True) strength=models.FloatField('Strength', max_digits=3, decimal_places=2, core=True)
Trying to add a Node through the Admin gives this traceback (SVN r5061):
Traceback (most recent call last): File "C:\cygwin\home\Dan\src\django_src\django\template\__init__.py" in render_node 723. result = node.render(context) File "C:\cygwin\home\Dan\src\django_src\django\template\defaulttags.py" in render 125. nodelist.append(node.render(context)) File "C:\cygwin\home\Dan\src\django_src\django\contrib\admin\templatetags\admin_modify.py" in render 170. bound_related_object = relation.bind(context['form'], original, bound_related_object_class) File "C:\cygwin\home\Dan\src\django_src\django\db\models\related.py" in bind 129. return bound_related_object_class(self, field_mapping, original) TypeError at /admin/ntwk/node/add/ 'bool' object is not callable
Change History (6)
comment:1 Changed 10 years ago by
comment:2 Changed 10 years ago by
comment:3 Changed 9 years ago by
comment:4 Changed 9 years ago by
comment:5 Changed 9 years ago by
comment:6 Changed 9 years ago by
As expected, this is fixed with the merge of newforms-admin. I verified using my own models (I had to specify the 'fk_name' attribute on my inline class), and everything saves without error.
I am marking this as "worksforme" rather than "fixed" since I'm not the original reporter. Dan, feel free to move this ticket to "fixed", or alternatively the developers are welcome to do so as well.
Thanks!
This is the same bug as #4490. | https://code.djangoproject.com/ticket/4126 | CC-MAIN-2017-13 | en | refinedweb |
I posted a tutorial on my favorite programming forum </dream.in.code>, and thought Id go ahead and share it on my blog as well. One question I get all the time in programming communities, always by young, new programmers, is how to work with web Services in .Net. It was these questions that lead me to writing the tutorial I posted on Dream.In.Code.
I guess before you can show someone how to create and consume a Web Service, you need to ensure they know and understand what a Web Service actually is. The following is the easiest, simplest definition I've been able to come up for "What is a web service":
What is a Web Service you say? Well a Web Service is very general model for building applications that can be implemented for any operating system that supports communication over the web. To some this may sound a lot like a Web Site, but that is not the case, there are many differences between the two.
Here are the main differences between a Web Service and a Web Site:
So as you can see, a Web Service and a Web Application have almost the same role, they just go about fulfilling that role in vastly different ways. A Web Service uses HTTP (Hypertext Transfer Protocol) and SOAP (Simple Object Access Protocol) to transfer data to and from the clients. Data sent to and from the Web Service is rendered into XML so the Web Service can read it, then it is sent back to the slient, which then renders the returned XML into the format it was expecting.
First, lets take a look at creating a simple Web Service. To do this, open Visual Studio, once it loads click File > New Website, when the New Project Dialog opens, select Web Service from the list of available projects, then for Location select either HTTP or File System (your own IIS), and lastly for Language select C#.
When the project is created it creates a Service.amsx and Sercive.cs file, you can either rename these, or delete and add new ones with the name of SampleService asmx and SampleService.cs.Now double-click on SampleService.cs to open the code file, then make sure you have the following references at the top of your class:
using System;
using System.Web;
using System.Web.Services;
using System.Web.Services.Protocols;
using System.Data.SqlClient;
using System.Data;
using System.Configuration;
If all the references aren't there, add the ones that are missing. Next, you will notice the first 2 lines are out of the ordinary, if you've never created a Web Service these two lines are downright weird:
[WebService(Namespace = "")]
[WebServiceBinding(ConformsTo = WsiProfiles.BasicProfile1_1)]
Heres whats going on here, line 1 declares the Web Service's Namespace, and you might now be wondering Well whats a namespace? Well luckily theres an answer to that: of the first section of the Web Service uses the WebServiceBinding Method to create the Web Service's version of an Interface. After those 2 lines then you have your Initializer, which is, of course, the same as any other class. Here you can instantiate any objects you need used in each instance of the service:
public SampleService ()
{
//Uncomment the following line if using designed components
//InitializeComponent();
}
Nothing new there, now we get to jump into the heart of a Web Service, which is, of course, to make it do something, to perform some service. When you create your methods, if you want them to be visible from the client thats consuming the service, you have to add the [WebMethod()] Attribute to it. This tells the service that its ok to expose that method to the client which is consuming it.
Since this is just a simple, basic Web Service it only has a single method, just to show you what can be done with a Web Service. Since this service is communicating with a Sql Database, it needs access to the System.Data.SqlClient Namespace. There are other Namespaces the service will use, we listed them above, but lets take a closer look at them:
Since the is an example for a beginner with Web Services, our sample service has but a single method in it, GetAllUserInfo. This method is designed to:
Now lets take a look at the code for this);
}
}
This operates as explained. It opens a connection to our database, retrieves all the user information from the users table, then populates a Dataset ans returns the DataSet to the client. Notice the [WebMethod()] attribute at the start of the method, this says that any client that consumes this Web Service will have access to that method, and the data it returns.
Believe it or not, our Web Service is complete. Now with a Web Service in a production environment will have much more meat on its bones, but for our demonstration, this is really all we need to show how a Web Service works. Next question people ask is "Ok, we have this service out there on our server, how do we access it and use it?". Well thats a good question, to use the Web Service we just created you need a consumer, a client that will connect to the service and gather the data.
In this example we will create an ASP.Net Web Application that will consume our Web Service, then display the returned data in a GridView Control. So, in your Solution Explorer window, right-click on the very top node, the solution name, then select Add > New Project. Once the New Project Dialog opens, you will this time select ASP.Net Web Site as the project type, once again either HTTP or File System for the Location and C# for the Language.
When the project is created is automatically adds a Default.aspx, and Default.aspx.cs. It's the CS file we will be doing the programming in. You always want to keep your presentation seperate Default.aspx.cs, and make sure you have the following references at the top of your class:
using System;
using System.Data;
using System.Configuration;
using System.Collections;
using System.Web;
using System.Web.Security;
using System.Web.UI;
using System.Web.UI.WebControls;
using System.Web.UI.WebControls.WebParts;
using System.Web.UI.HtmlControls;
Those are the only references you will need for this example. Now we need to add a reference to the Web Service we just created, to do this:
NOTE: Before clicking the Add Reference button I renamed my from localhost to Sample
Now we have a reference to our Web Service and can starting using it. In our ASPX page we have a single method, BindGrid(), and this method will:
The code for this method();
}
Simple enough isnt it? Now to use this, we will call this method from our Page_Load Event in our page, and that looks like this:
protected void Page_Load(object sender, EventArgs e)
{
BindGrid();
}
Believe it or not, thats it, thats all the code required to consume our Web Service and sisplay the returned data. The one difference in our BindGrid method is this:
WebServiceConsumer.Sample.SampleService sample = new WebServiceConsumer.Sample.SampleService();
That is where we create the reference to our Web Service, then we can use sample.WhateverMethodIsThere for utilizing the methods in the Web Service, instead of typing the long name out.
I am also providing a links to the Source Code for this example, all I ask is that you keep the header intact.
Thank you for reading, and I hope you found this helpful | http://geekswithblogs.net/PsychoCoder/archive/2007/09/30/intro_to_web_services.aspx | CC-MAIN-2017-13 | en | refinedweb |
Service Bus queues, topics, and subscriptions
Microsoft Azure Service Bus supports a set of cloud-based, message-oriented-middleware technologies including reliable message queuing and durable publish/subscribe messaging. These "brokered" messaging capabilities can be thought of as decoupled messaging features that support publish-subscribe, temporal decoupling, and load balancing scenarios using the Service Bus messaging fabric. Decoupled communication has many advantages; for example, clients and servers can connect as needed and perform their operations in an asynchronous fashion.
The messaging entities that form the core of the messaging capabilities in Service Bus are queues, topics and subscriptions, and rules/actions..
A related benefit is "load leveling," which enables producers and consumers to send and receive messages at different rates. In many applications, the system load varies over time; however, the processing time required for each unit of work is typically constant. Intermediating message producers and consumers with a queue means that the consuming application only:
QueueDescription myQueue; myQueue = namespaceClient.CreateQueue("TestQueue"); MessagingFactory factory = MessagingFactory.Create(ServiceBusEnvironment.CreateServiceUri("sb", ServiceNamespace, string.Empty), credentials); QueueClient myQueueClient = factory.CreateQueueClient("TestQueue");
You can then send messages to the queue. For example, if you have a list of brokered messages called
MessageList, the code appears similar to the following:
for (int count = 0; count < 6; count++) { var issue = MessageList[count]; issue.Label = issue.Properties["IssueTitle"].ToString(); myQueueClient.Send(issue); }
You then Complete on the received message. When Service Bus sees the Complete call, it marks the message as being consumed.
If the application is unable to process the message for some reason, it can call the (essentially performing an Abandon operation by default)..
Topics and subscriptions.
By way of comparison, the message-sending functionality of a queue maps directly to a topic and its message-receiving functionality maps to a subscription. CreateTopic method. For example:
TopicDescription dataCollectionTopic = namespaceClient.CreateTopic("DataCollectionTopic");
Next, add subscriptions as desired:
SubscriptionDescription myAgentSubscription = namespaceClient.CreateSubscription(myTopic.Path, "Inventory"); SubscriptionDescription myAuditSubscription = namespaceClient.CreateSubscription(myTopic.Path, "Dashboard");
You can then create a topic client. For example:
MessagingFactory factory = MessagingFactory.Create(serviceUri, tokenProvider); TopicClient myTopicClient = factory.CreateTopicClient(myTopic.Path)
Using the message sender, you can send and receive messages to and from the topic, as shown in the previous section. For example:
foreach (BrokeredMessage message in messageList) { myTopicClient.Send(message); Console.WriteLine( string.Format("Message sent: Id = {0}, Body = {1}", message.MessageId, message.GetBody<string>())); }>())); }
Rules and actions:
namespaceManager.CreateSubscription("IssueTrackingTopic", "Dashboard", new SqlFilter("StoreName = 'Store1'"));
With this subscription filter in place, only messages that have the
StoreName property set to
Store1 are copied to the virtual queue for the
Dashboard subscription.
For more information about possible filter values, see the documentation for the SqlFilter and SqlRuleAction classes. Also, see the Brokered Messaging: Advanced Filters and Topic Filters samples.
Next steps
See the following advanced topics for more information and examples of using Service Bus messaging. | https://docs.microsoft.com/en-us/azure/service-bus-messaging/service-bus-queues-topics-subscriptions | CC-MAIN-2017-13 | en | refinedweb |
NAME
exit - cause normal process termination
SYNOPSIS
#include <stdlib.h> void exit(int status);
DESCRIPTION
The exit() function causes normal process termination and the value of status & 0377 VALUE
The exit() function does not return.
ATTRIBUTES
Multithreading (see pthreads(7)) The exit() function uses a global variable that is not protected, so it is not thread-safe.
CONFORMING TO
SVr4, 4.3BSD, POSIX.1-2001, C89, C99.
NOTES
It is undefined what happens;.74 of the Linux man-pages project. A description of the project, information about reporting bugs, and the latest version of this page, can be found at. | http://manpages.ubuntu.com/manpages/wily/en/man3/exit.3.html | CC-MAIN-2017-13 | en | refinedweb |
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Scaling PostgreSQL Performance Using Table Partitioning
Here’s a classic scenario
You work on a project that stores transactional data in a database. The application gets deployed to production and early on the performance is great, selecting data from the database is snappy and insert latency goes unnoticed. Over a time period of days/weeks/months the database starts to get bigger and queries slow down.
There are various approaches that can help you make your application and database run faster. A Database Administrator (DBA) will take a look and see that the database is tuned. They offer suggestions to add certain indexes, move logging to separate disk partitions, adjust database engine parameters and verify that the database is healthy. You can also add Provisioned iOPS on EBS volumes or obtain faster (not just separate) disk partitions. This will buy you more time and may resolve this issues to a degree.
At a certain point you realize the data in the database is the bottleneck.
In many application databases some of the data is historical information that becomes less important after a certain amount of time. If you figure out a way to get rid of this data your queries will run faster, backups run quicker and use less disk space. You can delete it but then the data is gone forever. You could run a slew of DELETE statements causing a ton of logging and consuming database engine resources. So how do we get rid of old data efficiently but not lose the data forever?
Table Partitioning
Table partitioning is a good solution to this very problem. You take one massive table and split it into many smaller tables - these smaller tables are called partitions or child tables. Operations such as backups, SELECTs, and DELETEs can be performed against individual partitions or against all of the partitions. Partitions can also be dropped or exported as a single transaction requiring minimal logging.
Terminology
Let’s start with some terminology that you will see in the remainder of this blog.
Master Table
Also referred to as a Master Partition Table, this table is the template child tables are created from. This is a normal table, but it doesn’t contain any data and requires a trigger (more on this later). There is a one-to-many relationship between a master table and child tables, that is to say that there is one master table and many child tables.
Child Table
These tables inherit their structure (in other words, their Data Definition Language__or __DDL__for short) from the master table and belong to a single master table. The child tables contain all of the data. These tables are also referred to as __Table Partitions.
Partition Function
A __partition function__is a Stored Procedure that determines which child table should accept a new record. The master table has a trigger which calls a partition function. There are two typical methodologies for routing records to child tables:
- By Date Values- An example of this is purchase order date. As purchase orders are added to the master table, this function is called by the trigger. If you create partitions by day, each child partition will represent all the purchase orders entered for a particular day. This method is covered by this article.
- By__Fixed Values__- An example of this is by geographic location such as states. In this case, you can have 50 child tables, one for each US state. As INSERTs are fired against the master the partition function sorts each new row into one of the child tables. This methodology isn’t covered by this article because it doesn’t help us remove older data.
Let’s Try Configuring Table Partitions!
The example solution demonstrates the following:
- Automatically creating database table partitions based on date
- Schedule the export of older table partitions to compressed flat files in the OS
- Drop the old table partition without impacting performance
- Reload older partitions so they are available to the master partition
Take the time to let that last piece of the solution sink in. Most of the documentation on partitioning I’ve read through simply uses partitioning to keep the database lean and mean. If you needed older data, you’d have to keep an old database backup. I’ll show you how you can keep the database lean and mean through partitioning but also have the data available to you when you need it without db backups.
Conventions
Commands run in shell as the root user will be prefixed by:
(root#)
Commands run in a shell as a non-root user, eg. postgres will be prefixed by:
postgres$
Commands run within the PostgreSQL database system will look as follows:
my_database>
What you’ll need
The examples below use PostgreSQL 9.2 on Engine Yard. You will also need git for installing plsh.
Summary of Steps
Here’s a summary of what we are going to do:
- Create a master table
- Create a trigger function
- Create a table trigger
- Create partition maintenance function
- Schedule the partition maintenance
- Reload old partitions as needed
Create a Master Table
For this example we’ll be creating a table to store basic performance data (cpu, memory, disk) about a group of servers (server_id) every minute (time).
CREATE TABLE myschema.server_master ( id BIGSERIAL NOT NULL, server_id BIGINT, cpu REAL, memory BIGINT, disk TEXT, "time" BIGINT, PRIMARY KEY (id) );
Notice that in the code above the column name time is in quotes. This is necessary because time is a keyword in PostgreSQL. For more information on Date/Time keywords and functions visit the PostgreSQL Manual.
Create Trigger Function
The trigger function below does the following
- Creates child partition child tables with dynamically generated “CREATE TABLE” statements if the child table does not already exist.
- Partitions (child tables) are determined by the values in the “time” column, creating one partition per calendar day.
- Time is stored in epoch format which is an integer representation of the number of seconds since 1970-01-01 00:00:00+00. More information on Epoch can be found at
- Each day has 86400 seconds, midnight for a particular day is an epoch date divisible by 86400 without a remainder.
- The name of each child table will be in the format of
myschema.server_YYYY-MM-DD.
CREATE OR REPLACE FUNCTION myschema.server_partition_function() RETURNS TRIGGER AS $BODY$ DECLARE _new_time int; _tablename text; _startdate text; _enddate text; _result record; BEGIN --Takes the current inbound "time" value and determines when midnight is for the given date _new_time := ((NEW."time"/86400)::int)*86400; _startdate := to_char(to_timestamp(_new_time), 'YYYY-MM-DD'); _tablename := 'server_'||_startdate; -- Check if the partition needed for the current record exists PERFORM 1 FROM pg_catalog.pg_class c JOIN pg_catalog.pg_namespace n ON n.oid = c.relnamespace WHERE c.relkind = 'r' AND c.relname = _tablename AND n.nspname = 'myschema'; -- If the partition needed does not yet exist, then we create it: -- Note that || is string concatenation (joining two strings to make one) IF NOT FOUND THEN _enddate:=_startdate::timestamp + INTERVAL '1 day'; EXECUTE 'CREATE TABLE myschema.' || quote_ident(_tablename) || ' ( CHECK ( "time" >= EXTRACT(EPOCH FROM DATE ' || quote_literal(_startdate) || ') AND "time" < EXTRACT(EPOCH FROM DATE ' || quote_literal(_enddate) || ') ) ) INHERITS (myschema.server_master)'; -- Table permissions are not inherited from the parent. -- If permissions change on the master be sure to change them on the child also. EXECUTE 'ALTER TABLE myschema.' || quote_ident(_tablename) || ' OWNER TO postgres'; EXECUTE 'GRANT ALL ON TABLE myschema.' || quote_ident(_tablename) || ' TO my_role'; -- Indexes are defined per child, so we assign a default index that uses the partition columns EXECUTE 'CREATE INDEX ' || quote_ident(_tablename||'_indx1') || ' ON myschema.' || quote_ident(_tablename) || ' (time, id)'; END IF; -- Insert the current record into the correct partition, which we are sure will now exist. EXECUTE 'INSERT INTO myschema.' || quote_ident(_tablename) || ' VALUES ($1.*)' USING NEW; RETURN NULL; END; $BODY$ LANGUAGE plpgsql;
Create a Table Trigger
Now that the Partition Function has been created an Insert Trigger needs to be added to the Master Table which will call the partition function when new records are inserted.
CREATE TRIGGER server_master_trigger BEFORE INSERT ON myschema.server_master FOR EACH ROW EXECUTE PROCEDURE myschema.server_partition_function();
At this point you can start inserting rows against the Master Table and see the rows being inserted into the correct child table.
Create Partition Maintenance Function
Now let’s put the master table on a diet. The function below was built generically to handle the partition maintenance, which is why you won’t see any direct syntax for server.
How it works
- All of the child tables for a particular master table are scanned looking for any partitions where the name of the partition corresponds to a date older than 15 days ago.
- Each “too old” partition is exported/dumped to the local file system by calling the db function
myschema.export_partition(text, text). More on this is in the next section.
- If and only if the export to the local filesystem was successful the child table is dropped.
- This function assumes that the folder
/db/partition_dumpexists on the local db server. More on this in the next section. If you are wondering where the partitions are exported to this is where you should look!
CREATE OR REPLACE FUNCTION myschema.partition_maintenance(in_tablename_prefix text, in_master_tablename text, in_asof date) RETURNS text AS $BODY$ DECLARE _result record; _current_time_without_special_characters text; _out_filename text; _return_message text; return_message text; BEGIN -- Get the current date in YYYYMMDD_HHMMSS.ssssss format _current_time_without_special_characters := REPLACE(REPLACE(REPLACE(NOW()::TIMESTAMP WITHOUT TIME ZONE::TEXT, '-', ''), ':', ''), ' ', '_'); -- Initialize the return_message to empty to indicate no errors hit _return_message := ''; --Validate input to function IF in_tablename_prefix IS NULL THEN RETURN 'Child table name prefix must be provided'::text; ELSIF in_master_tablename IS NULL THEN RETURN 'Master table name must be provided'::text; ELSIF in_asof IS NULL THEN RETURN 'You must provide the as-of date, NOW() is the typical value'; END IF; FOR _result IN SELECT * FROM pg_tables WHERE schemaname='myschema' LOOP IF POSITION(in_tablename_prefix in _result.tablename) > 0 AND char_length(substring(_result.tablename from '[0-9-]*$')) <> 0 AND (in_asof - interval '15 days') > to_timestamp(substring(_result.tablename from '[0-9-]*$'),'YYYY-MM-DD') THEN _out_filename := '/db/partition_dump/' || _result.tablename || '_' || _current_time_without_special_characters || '.sql.gz'; BEGIN -- Call function export_partition(child_table text) to export the file PERFORM myschema.export_partition(_result.tablename::text, _out_filename::text); -- If the export was successful drop the child partition EXECUTE 'DROP TABLE myschema.' || quote_ident(_result.tablename); _return_message := return_message || 'Dumped table: ' || _result.tablename::text || ', '; RAISE NOTICE 'Dumped table %', _result.tablename::text; EXCEPTION WHEN OTHERS THEN _return_message := return_message || 'ERROR dumping table: ' || _result.tablename::text || ', '; RAISE NOTICE 'ERROR DUMPING %', _result.tablename::text; END; END IF; END LOOP; RETURN _return_message || 'Done'::text; END; $BODY$ LANGUAGE plpgsql VOLATILE COST 100; ALTER FUNCTION myschema.partition_maintenance(text, text, date) OWNER TO postgres; GRANT EXECUTE ON FUNCTION myschema.partition_maintenance(text, text, date) TO postgres; GRANT EXECUTE ON FUNCTION myschema.partition_maintenance(text, text, date) TO my_role; The function below is again generic and allows you to pass in the table name of the file you would like to export to the operating system and the name of the compressed file that will contain the exported table. -- Helper Function for partition maintenance CREATE OR REPLACE FUNCTION myschema.export_partition(text, text) RETURNS text AS $BASH$ #!/bin/bash tablename=${1} filename=${2} # NOTE: pg_dump must be available in the path. pg_dump -U postgres -t myschema."${tablename}" my_database| gzip -c > ${filename} ; $BASH$ LANGUAGE plsh; ALTER FUNCTION myschema.export_partition(text, text) OWNER TO postgres; GRANT EXECUTE ON FUNCTION myschema.export_partition(text, text) TO postgres; GRANT EXECUTE ON FUNCTION myschema.export_partition(text, text) TO my_role;
Note that the code above uses the
plsh language extension which is explained below. Also note that on our systems bash is located at /bin/bash, this may vary.
That Was Fun, Where are we?
Almost there. So far we’ve made all the necessary changes within the database to accommodate table partitions:
- Created a new master table
- Created the trigger and trigger function for the master table
- Created partition maintenance functions to export older partitions to the os and drop the old partitions
You could stop here if you’d like and proceed to the section “Now Let’s See it in Action” but sure to continue below to configure automated maintenance.
What we have left to do for automated maintenance is:
- Install the plsh extension
- Setup the os to store partition dumps
- Create a cron job to automate the calling of the maintenance partition function
Configure PostgreSQL and OS
Enabling PLSH in PostgreSQL
The PLSH extension is needed for PostgreSQL to execute shell commands. This is used by
myschema.export_partition(text,text) to dynamically create a shell string to execute
pg_dump. Starting as root, execute the following commands,
root# cd /usr/local/src # Build the extension .so files for postgresql root# curl -L -o plsh.tar.gz root# tar zxf plsh.tar.gz root# cd plsh-*/ root# make all install # Note that the postgres header files must be available root# su - postgres # Or whatever account postgresql is running under postgres$ psql my_database # Substitute the name of your database with the partitioned tables my_database> CREATE EXTENSION plsh; # NOTE: This must be done once for each database
Create the directory,
root# mkdir -p /db/partition_dump
Ensure that the postgres user owns the directory, and your deployment user’s group has permissions as well so that it can read the files. The default deploy user is ‘deploy’ on Engine Yard Cloud.
root# chown postgres:deploy /db/partition_dump
Even further information on PL/SH can be found in the plsh project’s README.
Schedule Partition Maintenance
The commands below will schedule the
partition_maintenance job to run at midnight every day
root# su - postgres ## This is the OS user that will run the cron job postgres$ mkdir -p $HOME/bin/pg_jobs ## This is the folder that will contain the script below postgres$ cat > $HOME/bin/pg_jobs/myschema_partition_maintenance #!/bin/bash # NOTE: psql must be available in the path. psql -U postgres glimpse <<SQL SELECT myschema.partition_maintenance('server'::text, 'server_master'::text, now()::date ); SQL ## Now press the <ctrl+d> keys to terminate the “cat” commands input mode postgres$ exit ## Exit the postgres os user and execute as root: root# chmod +x /home/postgres/bin/pg_jobs/myschema_partition_maintenance # Make script executable< root# crontab -u postgres -e ## Add the line: 0 0 * * * /home/postgres/bin/pg_jobs/myschema_partition_maintenance
View the cron jobs for the postgres user to ensure the crontab line is correct:
root# crontab -u postgres -l 0 0 * * * /home/postgres/bin/pg_jobs/myschema_partition_maintenance
Make sure your
/db/partition_dump folder is backed up if you are not using an Engine Yard Cloud instance. If you ever need the data again you’ll need these files to restore the old partitions. This may be as simple as rsyncing (copying) these files to another server just to be sure. We find that sending these to S3 works well for archival purposes.
Now your master tables are scheduled for partition maintenance and you can rest knowing that you’ve create something special; a nimble database that will keep itself on a weight loss program!
Reload Old Partitions
If you have separation anxiety from your old data or maybe a dull compliance request landed on your desk then you can reload the old partitions from the file system.
To reload a partition first we navigate to
/db/partition_dump on the local db server and identify the file and then as the postgres user we import the file back into the database.
postgres$ cd /db/partition_dump postgres$ ls # find the filename of the partition dump you need postgres$ psql my_database < name_of_partition_dump_file
After the partition file is loaded it will be queryable from the master table. Be aware that when the next time the partition maintenance job runs the newly imported partition will be exported again.
Now Let’s See it in Action
Create Child Tables
Let’s insert two rows of data to see creation of new child partitions in action. Open a psql session and execute the following:
postgres$ psql my_database my_database> INSERT INTO myschema.server_master (server_id, cpu, memory, disk, time) VALUES (123, 20.14, 4086000, '{sda1:510000}', 1359457620); --Will create "myschema"."servers_2013-01-29" my_database> INSERT INTO myschema.server_master (server_id, cpu, memory, disk, time) VALUES (123, 50.25, 4086000, '{sda1:500000}', 1359547500); --Will create "myschema"."servers_2013-01-30"
So what happened? Assuming this is the first time you’ve run this two new child tables were created, see the comments inline with the sql statement on the child tables that were created. The first insert can be seen by selecting against either the parent or child:
SELECT * FROM myschema.server_master; --Both records seen SELECT * FROM myschema."server_2013-01-29"; --Only the first insert is seen
Note the use of double quotes around the child partition table name, they aren’t there because the table is inherited, they are there because of hyphen used between the year-month-day.
Perform Partition Maintenance
The two rows we inserted are more than 15 days old. Manually running the partition maintenance job (same job as would be run by cron) will export these two partitions to the os and drop the partitions.
postgres$ /home/postgres/bin/pg_jobs/myschema_partition_maintenance
When the job is done you can see the two exported files:
postgres$ cd /db/partition_dump postgres$ ls -alh … -rw------- 1 postgres postgres 1.0K Feb 16 00:00 servers_2013-01-29_20130216_000000.000000.sql.gz -rw------- 1 postgres postgres 1.0K Feb 16 00:00 servers_2013-01-30_20130216_000000.000000.sql.gz
Selecting against the master table should yield 0 rows, the two child tables will also no longer exist.
Reload Exported Partitions
If you want to reload the first child partition from the exported file, gunzip it then reload it using psql:
postgres$ cd /db/partition_dump postgres$ gunzip servers_2013-01-29_20130216_000000.000000.sql.gz postgres$ psql my_database < servers_2013-01-29_20130216_000000.000000.sql
Selecting against the master table will yield 1 row, the first child table will now exist as well.
Notes
Our database files reside on a partition mounted at /db which is separate from our root (‘/’) partition.
For more information on PostgreSQL extensions visit the extensions documentation. The database engine doesn’t return the number of rows affected correctly (always 0 rows affected) when performing INSERTs and UPDATEs against the master table. If you use Ruby, be sure to adjust your code for the fact that the pg gem won’t have the correct value when reporting cmd_tuples. If you are using an ORM then hopefully they adjust for this accordingly.
Make sure you are backing up the exported partition files in
/db/partition_dump, these files lie outside of the database backup path.
The database user that is performing the INSERT against the master table needs to also have DDL permissions to create the child tables.
There is a small performance impact when performing an INSERT against a master table since the trigger function will be executed.
Ensure that you are running the absolute latest point release for your version of PostgreSQL, this ensures you are running the most stable and secure version available.
This solution works for my situation, your requirements may vary so feel free to modify, extend, mutilate, laugh hysterically or copy this for your own use.
Next Steps
One of the original assumptions was the creation of partitions for each 24 hour period, but this can be any interval of time (1 hour, 1 day, 1 week, every 435 seconds) with a few modifications. Part II of this blog post will discuss the necessary changes needed to the partition_maintenance function and table trigger. I’ll also explore how to create a second “archive” database that you can use to automatically load old partition data, keeping the primary database lean and mean for everyday use.
Share your thoughts with @engineyard on Twitter | http://blog.engineyard.com/2013/scaling-postgresql-performance-table-partitioning | CC-MAIN-2017-13 | en | refinedweb |
vxtrace(7) VxVM 3.5 vxtrace(7)
1 Jun 2002
NAME
vxtrace - VERITAS Volume Manager I/O Tracing Device
SYNOPSIS
/dev/vx/trace
DESCRIPTION
The vxtrace device implements the VERITAS Volume Manager (VxVM) I/O
tracing and the error tracing. An I/O tracing interface is available
that users or processes can use to get a trace of I/Os for specified
sets of kernel objects. Each separate user of the I/O tracing
interface can specify the set of desired trace data independent of all
other users. I/O events include regular read and write operations,
special I/O operations (ioctls), as well as special recovery
operations (for example, recovery reads). A special tracing mechanism
exists for getting error trace data. The error tracing mechanism is
independent of any I/O tracing and is always enabled for all pertinent
kernel I/O objects. It is possible for a process to get both a set of
saved errors and to wait for new errors.
IOCTLS
The format for calling each ioctl command is:
#include <<<<sys/types.h>>>>
#include <<<<vxvm/voltrace.h>>>>
struct tag arg;
int ioctl (int fd, int cmd, struct tag arg);
The first argument fd is a file descriptor which is returned from
opening the /dev/vx/trace device. Each tracing device opened is a
cloned device which can be used as a private kernel trace channel.
The value of cmd is the ioctl command code, and arg is usually a
pointer to a structure containing the arguments that need to be passed
to the kernel.
The return value for all these ioctls is 0 if the command was
successful, and -1 if it was rejected. If the return value is -1,
errno is set to indicate the cause of the error.
The following ioctl commands are supported:
VOLIOT_ERROR_TRACE_INIT
This command accepts no argument. The VOLIOT_ERROR_TRACE ioctl
initializes a kernel trace channel to return error trace data.
The trace channel will be initialized to return any previously
accumulated error trace data that has not yet been discarded. The
accumulated trace data can be skipped by issuing VOLIOT_DISCARD
on the channel. This call can be issued on a trace channel that
was previously initialized either for error tracing or for
regular I/O tracing. In this case, the channel is effectively
closed down and then reinitialized as described above. To get
- 1 - Formatted: August 2, 2006
vxtrace(7) VxVM 3.5 vxtrace(7)
1 Jun 2002
the error trace data, issue the read(2) system call. The error
trace data consists of a set of variable length trace event
records. The first byte of each record indicates the length, in
bytes, of the entire record (including the length byte), the
second byte indicates the type of the entry (which can be used to
determine the format of the entry). Each call to read() returns
an integral number of trace event records, not to exceed the
number of bytes requested in the read() call; the return value
from read() will be adjusted to the number of bytes of trace data
actually returned. If the O_NONBLOCK flag is set on the trace
channel, and no trace data is available, EAGAIN will be returned;
otherwise, the read will block interruptibly until at least one
trace record is available. When some trace data is available, the
available unread trace records will be returned, up to the limit
specified in the call to read(). If more trace records are
available, subsequent reads will return those records.
VOLIOT_IO_TRACE_INIT
The VOLIOT_IO_TRACE_INIT ioctl initializes a kernel trace channel
to return I/O trace data. This command accepts bufsize as the
argument. Initially, no objects are selected for I/O tracing. To
select objects to trace, issue the VOLIOT_IO_TRACE ioctl. The
bufsize argument specifies the kernel buffer size to use for
gathering events. A larger size reduces the chance that events
are lost due to scheduling delays in the event reading process. A
bufsize value of 0 requests a default size which is considered
reasonable for the system. The value of bufsize will be silently
truncated to a maximum value to avoid extreme use of system
resources. A bufsize value of (size_t)-1 will yield the maximum
buffer size.
VOLIOT_IO_TRACE,VOLIOT_IO_UNTRACE
The VOLIOT_IO_TRACE and VOLIOT_IO_UNTRACE ioctls enable and
disable, respectively, I/O tracing for particular sets of objects
on an I/O tracing channel. They both accept a voliot_want_list
structure tracelist as the argument. The tracelist argument
specifies object sets. The voliot_want_list structure specifies
an array of desired object sets. Each object set is identified by
a union of structures (the voliot_want_set union), each
representing different types of object sets. See the declaration
of these structures in voltrace.h for more detail.
FILES
/dev/vx/trace
SEE ALSO
vxintro(1M), vxtrace(1M), vxvol(1M) ioctl(2), read(2), vxconfig(7),
vxiod(7)
- 2 - Formatted: August 2, 2006 | http://modman.unixdev.net/?sektion=7&page=vxtrace&manpath=HP-UX-11.11 | CC-MAIN-2017-13 | en | refinedweb |
Coffeehouse Thread5 posts
Grrrr. Giving with one hand, taking with the other
Back to Forum: Coffeehouse
Comments have been closed since this content was published more than 30 days ago, but if you'd like to continue the conversation, please create a new thread in our Forums, or Contact Us and let us know.
<rant>>
After googling a bit I found this
I'm losing track how many times CodeProject.com has saved me from hours of annoying code!
.NET does actually let you write XOR'd lines in managed code, but it doesn't call them XOR and they aren't found in the standard Graphics routines. They are found in the System.Forms namespace in the ControlPaint class.
The ControlPaint class has a bunch of useful methods for building controls and hit has methods for doing XOR lines and boxes:
[Visual Basic]
Public Shared Sub DrawReversibleFrame( _
ByVal rectangle As Rectangle, _
ByVal backColor As Color, _
ByVal style As FrameStyle _
)
[Visual Basic]
Public Shared Sub DrawReversibleLine( _
ByVal start As Point, _
ByVal end As Point, _
ByVal backColor As Color _
)
[Visual Basic]
Public Shared Sub FillReversibleRectangle( _
ByVal rectangle As Rectangle, _
ByVal backColor As Color _
)
Hope this helps. I've noticed that the results of the methods in the ControlPaint class are similar to the styles that the VS.NET Forms Painter uses to show resize handles, selected items, etc.
ControlPaint Members
OK those are fine if you only want rectangles or lines.
However, they require having a control to paint on. What if I'm double buffering stuff on a different image, etc.
I think we could do with the raster ops of the old GDI world.
"Giving with one hand, taking with the other"
Also known as:
"The exchange of valuable information"
haha | https://channel9.msdn.com/Forums/Coffeehouse/9996-Grrrr-Giving-with-one-hand-taking-with-the-other | CC-MAIN-2017-13 | en | refinedweb |
?
.
This pyscopg2 (PostgreSQL) error usually signifies that some previous database query was incorrect (e.g., you tried to
order_by() a field that doesn't exist, or put a string in an integer column, etc.). That previous error aborted the transaction, causing all subsequent database access to fail with this message.
If you get this while at a shell, you can fix your database connection by executing a rollback:
from django.db import connection connection.cursor().execute('rollback')
If you get this from a view, it probably means the immediately previous query had a problem (but was caught by an over-eager exception handler).
In certain situations with PostgreSQL, a bogus error message about SET TIME ZONE may be returned. See #3179 (which is closed, but has a description of the problem). The real error message can probably be found in the postgres log file.") This is covered by #12285.
#15126 describes a common class of problem -- users with a single-item tuple forgetting the trailing comma. This was reported in the context of ModelForm.fields, but is a general problem that can (and should) be caught and reported..shortcuts
Invoking .render() with a <coconut> object instead of a HttpRequest as the first argument invokes AttributeError with the message: "'<coconut>' object has no attribute 'META'". Better will be something like, "render() takes a request object for the first argument."
django.template
- See this thread on the django-developers list for discussion of suppression of TypeError in the template engine
django.template.__init__.py
Using an invalid template tag results in the generic 'list index out of range' error which doesn't show any information about the offending file or expression. E.g., if a template contains:
... {% invalid_template_tag %} ...
An error occurs at line 279 of django.template.__init__.py:
278 try: 279 compiled_result = compile_func(self, token) 280 except KeyError: 281 self.invalid_block_tag(token, command) 282 except TemplateSyntaxError, e: 283 if not self.compile_function_error(token,e): 291 raise
What seems to be happening is that self.invalid_block_tag(...) raises an error which gets caught by TemplateSyntaxError and which eventually leads to the generic 'list index out of range' error being reported in the Django error page. | https://code.djangoproject.com/wiki/BetterErrorMessages?version=42 | CC-MAIN-2017-13 | en | refinedweb |
The Ember team has does an excellent job giving proper names to most their components, tools and libraries. For example, the rendering engine is called Glimmer, while it uses HTMLBars as the template language. Singletons in Ember applications are called Services. The build tool is called Ember-CLI, and a external application module is called an Addon, generally stored in NPM with the prefix
ember-cli-[addon-name]. Having recognizable names makes talking about them a lot easier.
This is very intentional for the community. There are specific terms for developers to discuss and an easy way to market changes to the framework. The latest is Engines, or Ember Engines.
The Ember Engines RFC started in October 2014 and was merged in April 2016. The goal was to allow large Ember applications to be split into consumable Addons allowing development teams to build logically-grouped pieces of an application in separate repositories and mount each micro application as a route or container in a parent application. In the table below are links to resources for EmberEngines for more history and details:
Engines and the flow to setup Engines were added to Ember fairly early in the Engines-RFC process. The most recent feature to be added, and what I think to be a crucial piece, is lazy-loading. This allows the core application to load with the initial page request while mounted engine sections of the application will be requested as needed in separate
.js files.
For applications that have sections with different business concerns, engines provide a structure for scaling without the threat of exponential file size growth. From the image above, the admin section of the site will only be used by a select group of individuals maintaining the site content. Allowing these user to load a separate file will shed weight from the readers using the blog app. The benefit lies in the main app maintaining the session, styles and global components.
To achieve the separation of concerns with engines, there are two ways to create the sandbox needed for mounting engines: in-repo-engine and external repo Addon. In this post, we’ll walk through building a basic application that uses an
in-repo-engine and
lazy-loading. In a the next post in the series, you’ll learn about making an Ember Addon an engine in an external git repository. In the final post of the series, we’ll bring it all together with shared services and links.
ember new large-company-site cd large-company-site ember install ember-engines
This assumes you have ember-cli install (
npm install -g ember-cli). Also, it assumes you are running Ember@2.10
These commands will setup your umbrella application with the appropriate addon to mount engines. The next step is creating an engine to mount. We will start with the in-app engine. While in the
large-company-site directory add an engine with
ember generate:
ember g in-repo-engine in-app-blog
This has added a directory named “lib” and an app addon directory structure named for “in-app-blog”.
Using the blueprint
in-repo-engine, ember-cli has added all the appropriate files to create a new app structure with its own routes. Open
lib/in-app-blog/addon/routes.js to add new routes:
import buildRoutes from 'ember-engines/routes'; export default buildRoutes(function() { - // Define your engine's route map here + this.route('new'); + this.route('posts'); }); });
Once the routes are added in the engine’s
addon/routes.js file, it is time to create route and template files for each. For this simple example, add the route and template files for
new and
posts in the
addon/routes and
addon/templates directories.
The next step is to add some content to see the routes working between the parent app and engine. In the following code examples you’ll add simple headlines to each
.hbs file. The file name will be in italics above the code block.
lib/in-app-blog/addon/templates/application.hbs
<h1>Blog Engine App</h1> {{outlet}}
lib/in-app-blog/addon/templates/new.hbs
<h1>New Form</h1> <form> {{input type="text" value=title}} {{textarea value=post}} <form>
lib/in-app-blog/addon/templates/posts.hbs
<h1>All Posts</h1> <ul> <!-- will insert items here --> </ul>
Now, add an application template to the umbrella application:
app/templates/application.hbs
<h1>Large Company Umbrella</h1> {{outlet}}
Finally, add a path to the mount route for the engine in
app/route.js:
import Ember from 'ember'; import config from './config/environment'; const Router = Ember.Router.extend({ location: config.locationType, rootURL: config.rootURL }); Router.map(function() { - this.mount('in-app-blog'); + this.mount('in-app-blog', {path: "blog"}); }); export default Router;
At this point, the structure is in place to create new routes, templates, controllers, and components for the blog engine application. The last change you need to make is in the
lib/in-app-blog/index.js file. You will change the application to lazy-load the blog engine. Add the following to the
index.js file in the
lib/in-app-blog:
/\\\*jshint node:true\\\*/ var EngineAddon = require('ember-engines/lib/engine-addon'); module.exports = EngineAddon.extend({ name: 'in-app-blog', + lazyLoading: true, isDevelopingAddon: function() { return true; } });
In the terminal, run
ember s, and open your browser to the location
localhost:4200/blog/posts.
Using the Chrome browser and the Developer Tools, you can open the Network tab to see the network traffic. What you’ll see is multiple
.js files being loaded.
Highlighted in the developer console is
engine.js which is a separate file from
large-company-site.js. This is it, this is what we’ve been waiting for. You can now built the largest Ember site ever and separate concerns with engines and know your users will get script and style resources efficiently. The benefit to you and your team is huge—you don’t have to spend all of your time configuring the build process. That’s Ember’s gift to you.
This example will be on Github at. As the series continues, the GitHub repo will include tags/commits for each blog post.
In the next post of the series, you will create an external addon as an engine and link it to your project. The final post will add shared services and links to tie the application together. | https://www.bignerdranch.com/blog/is-your-ember-app-too-big-split-it-up-with-ember-engines/?utm_source=javascriptweekly&utm_medium=email | CC-MAIN-2017-13 | en | refinedweb |
Hiermenus Go Forth, XIX - DHTML Lab | 5
Hiermenus Go Forth, XIX:
Version 4.0.12 - The Complete Script (Full-Window)
HM_ChildSecondsVisible
There have been many requests for a timer variable similar to HM_TopSecondsVisible but for child menus.
Currently, if HM_ClickKill is false, the menu tree collapses immediately upon user mouseout. Only the top-level menu remains visible, depending on the value of HM_TopSecondsVisible.
Since such a parameter is easy to implement, I have included it in 4.0.12 and not waited for a major release.
HM now recognizes the HM_GL_ChildSecondsVisible and HM_PG_ChildSecondsVisible parameter variables. If neither of these parameters are declared, the value for HM_ChildSecondsVisible defaults to .3, that is 300 milliseconds.
Note that, like HM_TopSecondsVisible, HM_ChildSecondsVisible takes a "seconds" value. A conversion to milliseconds is performed internally by HM.
Now, if a user mouses out of a menu tree, and HM_ClickKill is false, the tree will not collapse immediately, but only after the time specified in HM_ChildSecondsVisible has elapsed. If, during that time, the user mouses back onto the menu tree, the timer is, of course, cancelled.
Better Centering
In Version 4.0.9, when we improved JavaScript expression handling for menu positioning, we provided a "courtesy" function, HM_f_CenterMenu(), to help authors center permanent top-level menus on the page.
HM_f_CenterMenu() is not part of HM, but just an example of the type of function one can create and call from the positioning parameters. We included it in HM_Loader.js, and many authors are using it.
Currently HM_f_CenterMenu() centers a top-level menu regardless of menu width and window width. That is, if the menu is wider than the window, the menu is still centered in the page, with the menu's left edge positioned outside the window to the left. This is not desired, as the standard
behavior of centered HTML elements is that they are centered as long as the window width allows. If the element is wider than the window, the element is left-aligned. HM_f_CenterMenu() has been modified to left-align the menu to a "minimum left pixel position" which you can specify.
Thanks to Adrien Constant for pointing out this behavior.
function HM_f_CenterMenu(topmenuid) { //4.0.12 var MinimumPixelLeft = 0; var TheMenu = HM_DOM ? document.getElementById(topmenuid) : HM_IE4 ? document.all(topmenuid) : eval("window." + topmenuid); var TheMenuWidth = HM_DOM ? parseInt(TheMenu.style.width) : HM_IE4 ? TheMenu.style.pixelWidth : TheMenu.clip.width; var TheWindowWidth = HM_IE ? document.body.clientWidth : window.innerWidth; //4.0.12 // return ((TheWindowWidth-TheMenuWidth) / 2); return Math.max(parseInt((TheWindowWidth-TheMenuWidth) / 2),MinimumPixelLeft); }
Improved JS Expression Handling in Item Parameters
Version 4.0.12 also improves the handling of JS expressions used to define the
- item_is_rollover,
- item_permanently_highlighted, and
- item_has_child
item array element parameters.
Files Changed in Version 4.0.12
- HM_ScriptNS4.js
- HM_ScriptIE4.js
- HM_ScriptDOM.js
You will need to overwrite previous versions of the above files to upgrade to 4.0.12.
HM_Loader.js has also been changed to include the HM_GL_ChildSecondsVisible parameter and the
On the next page, the sample page included in the download.
Created: June 12, 2001
Revised: June 12, 2001
URL: | http://www.webreference.com/dhtml/column55/4.html | CC-MAIN-2017-13 | en | refinedweb |
First Tryst with Domain Aspects
I am tired of looking at aspects for logging, tracing, auditing and profiling an application. With the new AspectJ 5 and Spring integration, you can do all sorts of DI and wiring on aspects using the most popular IoC container. Spring's own
@Transactionaland
@Configurableare great examples of AOP under the hoods. However, I always kept on asking Show me the Domain Aspects, since I always thought that in order to make aspects a first class citizen in modeling enterprise applications, it has to participate in the domain model.
In one of our applications, we had a strategy for price calculation which worked with the usual model of injecting the implementation through the Spring container.
<beans>
<bean id="defaultStrategy" class="org.dg.domain.DefaultPricing"/>
<bean id="priceCalculation" class="org.dg.domain.PriceCalculation">
<property name="strategy">
<ref bean="defaultStrategy"/>
</property>
</bean>
</beans>
Things worked like a charm, till in one of the deployments, the client came back with the demand for implementing strategy failovers. The default implementation will continue to work as the base case, while in the event of failures, we need to iterate over a collection of strategies till one gets back with a valid result. Being a one of a kind of request, we decided NOT to change the base class and the base logic of strategy selection. Instead we chose to have a non-invasive way of handling the client request by implementing pricing strategy alternatives through a domain level aspect.
public aspect CalculationStrategySelector {
private List<ICalculationStrategy> strategies;
public void setStrategies(List<ICalculationStrategy> strategies) {
this.strategies = strategies;
}
pointcut inCalculate(PriceCalculation calc)
: execution(* PriceCalculation.calculate(..)) && this(calc);
Object around(PriceCalculation calc)
: inCalculate(calc) {
int i = 0;
int maxRetryCount = strategies.size();
while (true) {
try {
return proceed(calc);
} catch (Exception ex) {
if ( i < maxRetryCount) {
calc.setStrategy(getAlternativeStrategy(i++));
} else {
// handle exceptions
}
}
}
}
private ICalculationStrategy getAlternativeStrategy(int index) {
return strategies.get(index);
}
}
And the options for the selector were configured in the configuration xml of Spring ..
<beans>
<bean id="strategySelector"
class="org.dg.domain.CalculationStrategySelector"
factory-
<property name="strategies">
<list>
<ref bean="customStrategy1"/>
<ref bean="customStrategy2"/>
</list>
</property>
</bean>
<bean id="customStrategy1" class="org.dg.domain.CustomCalculationStrategy1"/>
<bean id="customStrategy2" class="org.dg.domain.CustomCalculationStrategy2"/>
</beans>
The custom selectors kicked in only when the default strategy fails. Thanks to AOP, we could handle this problem completely non-invasively without any impact on existing codebase.
And you can hide your complexities too ..
Aspects provide a great vehicle to encapsulate many complexities out of your development team. While going through Brian Goetz's Java Concurrency In Practice, I found one snippet which can be used to test your code for concurrency. My development team has just been promoted to Java 5 features and not all of them are enlightened with the nuances of
java.util.concurrent. The best way that I could expose the services of this new utility was through an aspect.
The following snippet is a class TestHarness and is replicated shamelessly from JCIP ..
package org.dg.domain.concurrent;
import java.util.concurrent.CountDownLatch;
public class TestHarness {
public long timeTasks(int nThreads, final Runnable task)
throws InterruptedException {
final CountDownLatch startGate = new CountDownLatch(1);
final CountDownLatch endGate = new CountDownLatch(nThreads);
for(int i = 0; i < nThreads; ++i) {
Thread t = new Thread() {
public void run() {
try {
startGate.await();
try {
task.run();
} finally {
endGate.countDown();
}
} catch (InterruptedException ignored) {}
}
};
t.start();
}
long start = System.nanoTime();
startGate.countDown();
endGate.await();
long end = System.nanoTime();
return end - start;
}
}
My target was to allow my developers to write concurrency test codes as follows ..
public class Task implements Closure {
@Parallel(5) public void execute(Object arg0) {
// logic
// ..
}
}
The annotation
@Parallel(5)indicates that this method need to be run concurrently in 5 threads. The implementation of the annotation is trivial ..
import java.lang.annotation.ElementType;
import java.lang.annotation.Retention;
import java.lang.annotation.RetentionPolicy;
import java.lang.annotation.Target;
@Target(ElementType.METHOD)
@Retention(RetentionPolicy.RUNTIME)
public @interface Parallel {
int value();
}
The interesting part is the main aspect which implements the processing of the annotation in AspectJ 5 .. Note the join point matching based on annotations and the context exposure to get the number of threads to use for processing.
public aspect Concurrency {
pointcut parallelExecutionJoinPoint(final Parallel par) :
execution(@Parallel public void *.execute(..)) && @annotation(par);
void around(final Parallel par) : parallelExecutionJoinPoint(par) {
try {
long elapsed =
new TestHarness().timeTasks(par.value(),
new Runnable() {
public void run() {
proceed(par);
}
} );
System.out.println("elapsed time = " + elapsed);
} catch (InterruptedException ex) {
// ...
}
}
}
The above example shows how you can build some nifty tools, which your developers will love to use. You can shield them from all complexities of the implementation and provide them the great feature of using annotations from their client code. Under the hoods, of course, it is AspectJ doing all the bigwigs.
In some of the future postings, I will bring out many of my encounters with aspects. I think we are in for an aspect awakening and the very fact that it is being backed by Spring, will make it a double whammy !
2 comments:
Interesting post. BTW we used aspects for billing, tx mgmt, tracing state change of domain objects and to a certain extent even for persistence. I will probably elaborate what I mean in a later post during the Diwali break. BTW enjoy reading your posts. There is so much of congruency between our thoughts.
I would be very much interested to know about usage of domain aspects in real life applications. As u have mentioned, usage of aspects in the service layer for tracing, logging, tx management has been very much mainstream. Sometime back I had documented my experiences of using AOP to implement failover, which came out in InfoQ. But I am yet to see extensive usage of domain aspects in the context of DDD. Would love to hear your detailed post coming up in Diwali. BTW do u have any blog which I can bookmark ? | http://debasishg.blogspot.com/2006/10/aspect-days-are-here-again.html | CC-MAIN-2017-39 | en | refinedweb |
How to load a custom QQuickItem from inside a library so that it gets registered & updated like other QQuickItems in the application
I have a
MyQuickItemclass derived from
QQuickItemas below
// MyQuickItem.hpp class MyQuickItem : public QQuickItem { Q_OBJECT public: MyQuickItem(); virtual ~ MyQuickItem(); protected: QSGNode* updatePaintNode(QSGNode *, UpdatePaintNodeData *) override; };
Following is
MyQuickItem.qml.
import QtQuick 2.0 import MyQuickItem 1.0 Item { MyQuickItem { id: my_quick_item objectName: "MyQuickItemObject" visible: false } }
Point to be noted is that all of above in a separate static library. And the library has a
qrcwhich has
MyQuickItem.qmlin it. This library has access to the global
QQmlApplicationEngineobject of the app as well.
My question: How can I load
MyQuickItemfrom inside my library so that it gets registered with QML like the other
QQuickItems in app's
main.qml?
I am trying something around the following way from inside my library in a C++ method called after main.qml is loaded by the application:
MyQuickItem * myItem = new MyQuickItem(); myItem->setParent(m_qml_engine->parent()); myItem->setParentItem(qobject_cast<QQuickItem*>(m_qml_engine->parent())); QQmlEngine::setObjectOwnership(myItem, QQmlEngine::JavaScriptOwnership); myItem->setHeight(500); // But myItem is NULL here !!! myItem->setHeight(500); // But myItem is NULL here !!!
Firstly, I don't know how to link
QUrl(QStringLiteral("qrc:/qml/MyQuickItem.qml"))to
myItempointer.
Secondly, doing the above does not seem to load
MyQuickItemcorrectly as I don't get a call to
updatePaintNodewhich I have overridden. I need the
Qt/QMLwindow system to call my
MyQuickItem ::updatePaintNodeas I have important logic there.
So, How can I correctly load
MyQuickItemfrom inside my library so that it gets registered & updated like other
QQuickItems?
Hi,
just a wild guess since I am a noob. Have you tried to include the header file from your lib in your main.cpp? If not try that and look up qmlRegisterType.
@Sikarjan I tried including
MyLibrary/MyQuickItem.hpp/ do a
qmlRegisterType<MyQuickItem>("MyQuickItem", 1, 0, "MyQuickItem"). But that does not help. I think this only works if I was doing an
import MyQuickItem 1.0to embed
MyQuickIteminto the Application's
main.qmlor any of its child. Here
MyQuickItem.qmlis inside a library. So struggling to make it visible under the application's qml tree
- SGaist Lifetime Qt Champion
Hi and welcome to devnet,
Did you already saw the QML Modules chapter in Qt's documentation ?
@SGaist I saw this example. but that one is not deriving the class from a
QQuickItem. In my case everything is working if I define
MyQuickItemin
main.qml& call
qmlRegisterTypefrom main.cpp. All I want to do is, move the
MyQuickIteminto its own
MyQuickItem.qmlinside my library without the app defining it. The goal is that the app should get
MyQuickItemfrom the library.
Isn't this possible without going into the QML module technique?
By the way, I can create a
QQuickitemin the following way.
QQuickItem * dynamic_quick_item = new QQuickItem(); dynamic_quick_item->setObjectName("DynamicQuickItemObject"); dynamic_quick_item->setHeight(500); dynamic_quick_item->setWidth(500);
I also have access to qml_engine & everything in
main.cpp.
But my problem is that : How can I add this
dynamic_quick_itemto the children list of qml objects?
- SGaist Lifetime Qt Champion
Then isn't the Creating C++ Plugins for QML chapter what you are looking for ?
- Roby Brundle | https://forum.qt.io/topic/82606/how-to-load-a-custom-qquickitem-from-inside-a-library-so-that-it-gets-registered-updated-like-other-qquickitems-in-the-application | CC-MAIN-2017-39 | en | refinedweb |
SINGAPORE (ICIS)--Here are some of the top stories from ICIS Asia and the Middle east for the week ended 21 February 2014.?xml:namespace>
Focus: China benzene imports muted on low-priced domestic cargoes
China benzene import prices will remain at a standstill, following a second reduction in the local major’s list price in 2014
Focus: India resists March acetone price hike on ample import arrivals Acetone buyers in India are resisting a proposed $20-30/tonne price hike for March-shipping cargoes, expecting substantial volume of imported material to arrive in the country this month
Focus: Asia naphtha to get boost from firming blending economics
Asia's naphtha prices are likely to be bolstered by improving gasoline blending economics in the region, as well as in Europe, which could potentially soak up heavy deep-sea supply that has been flowing into the region
Focus: Asia SM finds support at $1,600/tonne after recent plunge
Spot styrene monomer (SM) prices in Asia appear to have found support at $1,600/tonne this week after a recent plunge, with some market players bent on covering their short positions
Focus: India's amines prices decline on higher imports, weak demand
India's domestic ethanolamines prices are likely to fall further amid high import volumes which are outpacing consumption
Spot expandable polystyrene (EPS) prices in Asia may stay under pressure in the near term, as sellers have been cutting offers to draw in buyers amid continued weakness in demand | https://www.icis.com/resources/news/2014/02/24/9755877/asia-top-stories-weekly-summary/ | CC-MAIN-2017-39 | en | refinedweb |
Your answer is one click away!
I am working on some WebApplication with SpringBoot MVC pattern. I have four maven projects (DAO project, REST project(there is SpringBoot class for starting application), SERVICE project and CLIENT project). This projects are connected through dependencies.
My problem is with CLIENT project. There I have WelcomeController which is look like this :
package com.itengine.scoretracker.client.controller; import org.springframework.stereotype.Controller; import org.springframework.web.bind.annotation.RequestMapping; @Controller public class WelcomeController { @RequestMapping("/") public String welcome(){ return "static/html/index.html"; } }
And my html's are on this path:
When I relocate my static folder from CLIENT project in REST project on same location, my WelcomeController see index.html and everything works fine.
So please, can somebody help me with this issue, i really need this html's in CLIENT project. I dont have experience with configuration xml's because i learned on course SpringBoot without those xml's.
My web.xml's are empty, they have only this:
<!DOCTYPE web-app PUBLIC "-//Sun Microsystems, Inc.//DTD Web Application 2.3//EN" "java.sun.com/dtd/web-app_2_3.dtd"; > <web-app> <display-name>Archetype Created Web Application</display-name> </web-app>
My main class is like this:
package com.itengine.scoretracker.rest.init; import org.springframework.boot.SpringApplication; import org.springframework.boot.autoconfigure.SpringBootApplication; import org.springframework.boot.builder.SpringApplicationBuilder; import org.springframework.boot.context.web.SpringBootServletInitializer; import org.springframework.boot.orm.jpa.EntityScan; impo
Now, like u say @Reimeus, I have this situation:
And that won't working. And what with removing / mapping??
If your
index.html is plain html, your controller is unnecessary.
Just put the
index.html in
src/main/resources/static.
That's all you need. Spring boot takes care of the rest.
You also don't need a web.xml. | http://www.devsplanet.com/question/35270562 | CC-MAIN-2017-39 | en | refinedweb |
Your answer is one click away!
well i cant find how do this, basically its a variable union with params, basic idea, (writed as function)
Ex1
union Some (int le) { int i[le]; float f[le]; };
Ex2
union Some { int le; int i[le]; float f[le]; };
obs this don't works D: maybe a way to use an internal variable to set the lenght but don't works too. Thx.
No, this is not possible:
le would need to be known at compile-time.
One solution would be to use a templated union:
template <int N> union Some { int i[N]; float f[N]; };
N, of course, is compile-time evaluable.
Another solution is the arguably more succinct
typedef std::vector<std::pair<int, float>> Some;
or a similar solution based on
std::array.
Depending on your use case you could try to simulate a union.
struct Some { //Order is important private: char* pData; public: int* const i; float* const f; public: Some(size_t len) :pData(new char[sizeof(int) < sizeof(float) ? sizeof(float) : sizeof(int)]) ,i ((int*)pData) ,f ((float*)pData) { } ~Some() { delete[] pData; } Some(const Some&) = delete; Some& operator=(const Some&) = delete; };
Alternative solution using templates, unique_ptr and explicit casts:
//max_size_of<>: a recursive template struct to evaluate the // maximum value of the sizeof function of all types passed as // parameter //The recursion is done by using the "value" of another // specialization of max_size_of<> with less parameter types template <typename T, typename...Args> struct max_size_of { static const std::size_t value = std::max(sizeof(T), max_size_of<Args...>::value); }; //Specialication for max_size_of<> as recursion stop template <typename T> struct max_size_of<T> { static const std::size_t value = sizeof(T); }; //dataptr_auto_cast<>: a recursive template struct that // introduces a virtual function "char* const data_ptr()" // and an additional explicit cast operator for a pointer // of the first type. Due to the recursion a cast operator // for every type passed to the struct is created. //Attention: types are not allowed to be duplicate //The recursion is done by inheriting from of another // specialization of dataptr_auto_cast<> with less parameter types template <typename T, typename...Args> struct dataptr_auto_cast : public dataptr_auto_cast<Args...> { virtual char* const data_ptr() const = 0; //This is needed by the cast operator explicit operator T* const() const { return (T*)data_ptr(); } //make it explicit to avoid unwanted side e
C++ requires that the size of a type be known at compile time.
The size of a block of data need not be known, but all types have known sizes.
There are three ways around it.
I'll ignore the union part for now. Imagine if you wanted:
struct some (int how_many) { int data[how_many]; };
as the union part adds complexity which can be dealt with separately.
First, instead of storing the data as part of the type, you can store pointers/references/etc to the data.
struct some { std::vector<int> data; explicit some( size_t how_many ):data(how_many) {}; some( some&& ) = default; some& operator=( some&& ) = default; some( some const& ) = default; some& operator=( some const& ) = default; some() = default; ~some() = default; };
here we store the data in a
std::vector -- a dynamic array. We default copy/move/construct/destruct operations (explicitly -- because it makes it clearer), and the right thing happens.
Instead of a vector we can use a unique_ptr:
struct some { std::unique_ptr<int[]> data; explicit some( size_t how_many ):data(new int[how_many]) {}; some( some&& ) = default; some& operator=( some&& ) = default; some() = default; ~some() = default; };
this blocks copying of the structure, but the structure goes from being size of 3 pointers to being size of 1 in a typical
std implementation. We lose the ability to easily resize after the fact, and copy without writing the code ourselves.
The next approach is to template it.
template<std::size_t N> struct some { int data[N]; };
this, however, requires that the size of the structure be known at compile-time, and
some<2> and
some<3> are 'unrelated types' (barr
I would like to suggest a different approach: Instead of tying the number of elements to the
union, tie it outside:
union Some { int i; float f; }; Some *get_Some(int le) { return new Some[le]; }
Don't forget to
delete[] the return value of
get_Some... Or use smart pointers:
std::unique_ptr<Some[]> get_Some(int le) { return std::make_unique<Some[]>(le); }
You can even create a
Some_Manager:
struct Some_Manager { union Some { int i; float f; }; Some_Manager(int le) : m_le{le}, m_some{std::make_unique<Some[]>(le)} {} // ... getters and setters... int count() const { return m_le; } Some &operator[](int le) { return m_some[le]; } private: int m_le{}; std::unique_ptr<Some[]> m_some; };
Take a look at the Live example.
It's not possible to declare a structure with dynamic sizes as you are trying to do, the size must be specified at run time or you will have to use higher-level abstractions to manage a dynamic pool of memory at run time.
Also, in your second example, you include
le in the union. If what you were trying to do were possible, it would cause
le to overlap with the first value of
i and
f.
As was mentioned before, you could do this with templating if the size is known at compile time:
#include <cstdlib> template<size_t Sz> union U { int i[Sz]; float f[Sz]; }; int main() { U<30> u; u.i[0] = 0; u.f[1] = 1.0; }
If you want dynamic size, you're beginning to reach the realm where it would be better to use something like
std::vector.
#include <vector> #include <iostream> union U { int i; float f; }; int main() { std::vector<U> vec; vec.resize(32); vec[0].i = 0; vec[1].f = 42.0; // But there is no way to tell whether a given element is // supposed to be an int or a float: // vec[1] was populated via the 'f' option of the union: std::cout << "vec[1].i = " << vec[1].i << '\n'; } | http://www.devsplanet.com/question/35271453 | CC-MAIN-2017-39 | en | refinedweb |
.
Example: Correct Another Developer's Grammar!
Suppose you found a nice plugin to work with, but you realize that its owner doesn't speak English very well, and you see some poorly written text inside the code. Luckily the strings are translatable, so you're going to be able to change those strings with the help of the
gettext
If you don't want the word "the" in your slugs, you can delete them with the code snippet below:
<?php add_filter( 'sanitize_title', 'sanitize_title_example' ); function sanitize_title_example( $title ) { $title = str_replace( '-the-', '-', $title ); $title = preg_replace( '/^the-/', '', $title ); return $title; } ?>
A simple and elegant solution.
Setting Exceptions for Shortcode Texturization
This handy filter "allows you to specify which shortcodes should not be run through the
wptexturize() function", as said in the Codex.
Example: Exclude Your Shortcode from Texturization
If you want the shortcode you built to exempt from texturization, use this code to add your shortcode name to the "do not texturize" list:
<?php add_filter( 'no_texturize_shortcodes', 'no_texturize_shortcodes_example' ); function no_texturize_shortcodes_example( $shortcodes ) { $shortcodes[] = 'myshortcode'; return $shortcodes; } // Example source: ?>
Pretty easy, right?
Filtering a Comment's Approval Status
WordPress has its own checks for comments (which may be a little too easy against spammers) before deciding whether the comment should be marked as spam, be sent to the moderation queue or be approved. The
pre_comment_approve filter lets plugins help with this decision.
Example: Marking Comments With Long Author Names as Spam
In my country, Turkey, WordPress comment spammers usually use reeeaaally long names, sometimes the URL itself.
With the code snippet below, you can automatically eliminate spammers who use names like "Domestic and International Shipping With Extremely Low Prices (Click Here for More Information)":
<?php add_filter( 'pre_comment_approved', 'pre_comment_approved_example', 99, 2 ); function pre_comment_approved_example( $approved, $commentdata ) { return ( strlen( $commentdata['comment_author'] ) > 75 ) ? 'spam' : $approved; } // Example source: ?>
Special thanks to Andrew Norcross for the idea!
Bonus tip: If you want to eliminate spam by checking the length of the comment author's URL, use 'comment_author_url' instead of 'comment_author'. Andrew Norcross used the URL in his original tip, by the way.
Configuring the "Post By Email" Feature
Did you know that you can post to your WordPress blog via email? WordPress offers this seldom used feature and it allows you to turn it on or off with the
enable_post_by_email_configuration filter.
Example: Turning the "Post By Email" Feature On and Off
For some reason (such as security, maybe) you might want to turn off this feature. And you can do it with just one line of code:
<?php add_filter( 'enable_post_by_email_configuration', '__return_false', 100 ); ?>
Or if you're on WordPress Multisite and you need to enable this feature (since it's disabled by default on Multisite), you can use the
__return_true() function:
<?php add_filter( 'enable_post_by_email_configuration', '__return_true', 100 ); ?>
Filtering Your Page Titles
The
wp_title() function outputs the page titles, the ones that we see on our tab handles in browsers. And the wp_title function allows us to tamper with those titles.
Example: Rewriting Your Page Titles – The Right Way
A respected WordPress "guru" (and editor at Tuts+ Code) Tom McFarlin explains us in his blog how to rewrite our page titles properly with the
wp_title() function and the filter with the same name:
Since wp_title is a filtered function, this means that we're able to provide a custom hook that allows us to define the schema for displaying our titles not only more precisely, but also correctly.
<?php add_filter( 'wp_title', 'wp_title_example', 10, 2 ); function wp_title_example( = sprintf( __( 'Page %s', 'tuts_filter_example' ), max( $paged, $page ) ) . " $sep $title"; return $title; } // Example source: ?>
Be sure to check out his article. Thanks, Tom! you.
Example: Turn Down the Volume of Yellers
DO YOU GET A LOT OF COMMENTS IN WHICH EVERY SINGLE WORD IS UPPERCASE? If you do, you can automatically make those letters lowercase with the code snippet below:
<?php add_filter( 'preprocess_comment', 'preprocess_comment_example' ); function preprocess_comment_example( $commentdata ) { if( $commentdata['comment_content'] == strtoupper( $commentdata['comment_content'] )) $commentdata['comment_content'] = strtolower( $commentdata['comment_content'] ); return $commentdata; } // Example source: ?>
Cool, huh?
Managing Redirection After Login
This little filter allows us to set redirects (other than the administration panel) following the login process, which can be pretty useful in some cases.
Example: Redirect Subscribers to Website Home
If you don't want your users (with the role "Subscriber") to see your admin panel after they login, you can redirect them to your website's homepage:
<?php add_filter( 'login_redirect', 'login_redirect_example', 10, 3 ); function login_redirect_example( $redirect_to, $request, $user ) { global $user; if ( isset( $user->roles ) && is_array( $user->roles ) ) { if ( in_array( 'subscriber', $user->roles ) ) { return home_url(); } else { return $redirect_to; } } return; } ?>
The Codex warns us about one thing: "Make sure you use
add_filter outside
To add custom action links under your plugin's name at the list in the Plugin page, you can use this function and hook it to the filter:
<?php add_filter( 'plugin_action_links_' . plugin_basename( __FILE__ ), 'plugin_action_links_example' ); function plugin_action_links_example( $links ) { $links[] = '<a href="' . get_admin_url( null, 'options-general.php?page=my_plugin_settings' ) . '">' . __( 'Settings' ) . '</a>'; return $links; } // Example source: ?>
Notice that we're using the
__FILE__ constant to hook our function to the filter with your plugin's name. Neat, huh?
Use this with caution: If you abuse that area to fill there with links, people will remember you as a spammer.
Filtering the Content Inside the Post Editor
Ever wanted to pre-fill the post editor to start writing with a post template, or leaving notes for your authors? You can, thanks to the
the_editor_content filter.
Example: Leaving Reminders for Your Authors
Let's do the "leaving notes for authors" example: If you have a bunch of things to remind the writers of your blog, you can fill the post editor with HTML by using this code:
<?php add_filter( 'the_editor_content', 'the_editor_content_example' ); function the_editor_content_example( $content ) { // Only return the filtered content if it's empty if ( empty( $content ) ) { $template = 'Hey! Don\'t forget to...' . "\n\n"; $template .= '<ul><li>Come up with good tags for the post,</li><li>Set the publish time to 08:00 tomorrow morning,</li><li>Change the slug to a SEO-friendly slug,</li><li>And delete this text, hehe.</li></ul>' . "\n\n"; $template .= 'Bye!'; return $template; } else return $content; } // Example source: ?>
Change the
$template variable with anything you like and you're good to go!
End of Part Two
We went through the second
| https://code.tutsplus.com/tutorials/50-filters-of-wordpress-filters-11-20--cms-21296 | CC-MAIN-2017-39 | en | refinedweb |
Hi Guys,
I wanted to know how to access the methods of grandparent class from the child class (when those methods in child class are overridden) using super keyword. Please help.
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Hi Guys,
I wanted to know how to access the methods of grandparent class from the child class (when those methods in child class are overridden) using super keyword. Please help.
Why do you think you need to do this? Sounds like a symptom of bad design to me.
Just a thought, as in how do you get it.
I'm not convinced you can, and I'm not convinced you should. Like I said, it's more likely a symptom of a bad design.
I get what you are saying. But as I was attending a Java training class, a student came up with this question, so we were asked to find an answer for the same. As 'super' is used to get the methods of parent class, how can super be manipulated to get the methods of grandparent class.
Well, let me know what you come up with. I don't see an obvious way, other than adding methods in the parent class that can call methods in the grandparent class.
As Kevin mentioned, this is poor design and a situation that should be avoided at all costs - if something is designed this way, its time to redesign. That being said, you could use reflection to accomplish this, but its an ugly solution to an already ugly problem that may not always work.
I tried this:
Code java:
public class InnerClassTest { public static void main(String... args){ new ClassC().print(); } } class ClassA{ public void print(){ System.out.println("A"); } } class ClassB extends ClassA{ public void print(){ System.out.println("B"); } } class ClassC extends ClassB{ public void print(){ try { ((Class<ClassC>)this.getClass().getSuperclass().getSuperclass()).getMethod("print").invoke(this); } catch (Exception e) { e.printStackTrace(); } } }
But that just ends up calling ClassC's print() method, which results in a StackOverflowException, which is what I would expect. Hmph.
I think its a lot uglier than that. You need to get the grandparent Class, create a new instance of said class, - and this is where it gets ugly real fast - set all the appropriate values of the newly constructed grandparent object (assuming the classes are 'bean'-like, with getters and setters, you use the child values to set the grandparent values), then finally invoke the method. Did I mention this is ugly? ;)
This thread has been cross posted here:
Although cross posting is allowed, for everyone's benefit, please read:
Java Programming Forums Cross Posting Rules
The Problems With Cross Posting
oh I didnt know about that
Ack, crossposting, not cool. | http://www.javaprogrammingforums.com/%20java-theory-questions/11569-super-keyword-printingthethread.html | CC-MAIN-2017-39 | en | refinedweb |
Hi, i'm second year in college and i'm diving into the basics of c++ , basicly i have to read 2 numbers and divide them with each other by repeated substractions using a function, the function is to divide those numbers by repeatedly substracting one from another, and return the result, and the rest, + give an error in case one of the numbers is 0 :D
i tried to play with that a bit but i think i pretty much failed :(
here is what i got for now:
#include "stdafx.h" #include <conio.h> #include <iostream> using namespace std; int div (int x, int y) { int c=0,d=0; while (x!=0,y!=0,x>=y) { c=x-y; x=c; d=d+1; } return d; } int main() { int a,b,c=0,d=0,z=0,n; cout<<"A= "; cin>>a; cout<<" B= "; cin>>b; z = div (a,b); if (a==0 || b==0) { cout<<" ERROR "<<endl; } if (a!=b,a<=b) { cout<<" "<<endl; cout<<"A / B = "<<z<<endl; n=a%b; cout<<"Rest= "<<n<<endl; } getch(); }
i get 4 errors but it would be pointless to post them since i know its probably something obvious >.< | https://www.daniweb.com/programming/software-development/threads/250278/simple-problem-function-trouble | CC-MAIN-2017-39 | en | refinedweb |
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Thread: Global Variables In Java
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Global Variables In Java
How would I declare the variable numbers as global? I want to be able to access the same variable from any method so that the value of the variable is the same.
Code:
import TerminalIO.KeyboardReader; public class SortAndDestroy{ public static void main (String[] args){ int[] numbers = {42,37,5,33,6}; KeyboardReader reader = new KeyboardReader(); int choice; for(int i = 0; i < Names.numbers.length; i++){ System.out.print(numbers[i] + " "); } choice = reader.readInt("\nSelection[1] or Bubble[2] Sort? "); if (choice == 1) selectionSort(numbers); else if (choice == 2) bubbleSort(numbers); else System.out.println("That is not a choice."); } public static void selectionSort(int[] numbers){ for(int i = 0; i < numbers.length; i++){ int minIndex = findMinimum(numbers, i); if(minIndex != 1) swap(numbers, i, minIndex); System.out.print(numbers[i] + " "); } } public static int findMinimum(int[] numbers, int first){ int minIndex = first; for(int i = first + 1; i < numbers.length; i++) if(numbers[i] < numbers[minIndex]) minIndex = i; return minIndex; } public static void swap(int[] numbers, int x, int y){ int temp = numbers[x]; numbers[x] = numbers[y]; numbers[y] = temp; } public static void bubbleSort(int[] numbers){ int k = 0; boolean exchangeMade = true; while((k < numbers.length - 1) && exchangeMade){ exchangeMade = false; k++; for(int j = 0; j < numbers.length - k; j++) if(numbers[j] > numbers[j + 1]){ swap(numbers,j,j + 1); exchangeMade = true; } } } }
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public class SortAndDestroy{
int[] numbers = {42,37,5,33,6};
public static void main (String[] args){
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That's usually a bad idea. You don't want to give everyone access to a local variable. Instead make the variable private but create a public method within that class that returns the value.
Edit: Never mind. I misread your question. I thought you wanted to give other classes access to the variables in that class. As was posted already just declare the variable outside of your methods. That will put it into the scope of all your methods.
Last edited by Spookster; 05-27-2004 at 06:03 AM.Spookster
CodingForums Supreme Overlord
All Hail Spookster | http://www.codingforums.com/java-and-jsp/39335-global-variables-java.html | CC-MAIN-2017-39 | en | refinedweb |
def hotel_cost(nights): return 140 * nightsdef plane_ride_cost(city): if city == "Charlotte": return 183 elif city == "Tampa": return 220 elif city == "Pittsburgh": return 222 elif city == "Los Angeles": return 475def rental_car_cost(days): rents = 40 * days if days >= 7: return rents - 50 elif 7 > days >= 3: return rents - 20 else: return rents
def trip_cost(days,city): return hotel_cost(days) + plane_ride_cost(city) + rental_car_cost(days)
Oops, try again. trip_cost('Pittsburgh', 5) raised an error: cannot concatenate 'str' and 'NoneType' objects
HELP ,What's the matter?
I can't help you if I can't see your code.
could you see my code now?
This might be a syntax error if so, I'm going to need you to post your formatted code. If you don't know how to format code in your posts, please refer to:
oh dear my,it must be a bug.the codes must be def trip_cost(city,days):,not def trip_cost(days,city):.thx all the same.
It's not so much a bug, as it's just Codecademy's SCT being picky.
You are so nice,thank you!
This topic was automatically closed 7 days after the last reply. New replies are no longer allowed. | https://discuss.codecademy.com/t/pull-it-together/79729/4 | CC-MAIN-2017-39 | en | refinedweb |
Content-type: text/html
strpbrk - Find one of a set of bytes in a string
Standard C Library (libc.so, libc.a)
#include <string.h>
char *strpbrk(
const char *s1,
const char *s2);
Interfaces documented on this reference page conform to industry standards as follows:
strpbrk(): XPG4, XPG4-UNIX
Refer to the standards(5) reference page for more information about industry standards and associated tags.
Specifies a string being searched. Specifies a set of bytes in a string.
The strpbrk() function scans the string pointed to by the s1 parameter for the first occurrence of any byte in the string pointed to by the s2 parameter. The strpbrk() function treats the s2 parameter as a series of bytes and does not search for multibyte characters. The wcspbrk() function provides the same functionality but searches for characters rather than bytes.
Upon successful completion, the strpbrk() function returns a pointer to the matched byte. When no byte in the string pointed to by the s2 parameter occurs in the string pointed to by the s1 parameter, a null pointer is returned.
Functions: string(3), wcspbrk(3)
Standards: standards(5) delim off | https://backdrift.org/man/tru64/man3/strpbrk.3.html | CC-MAIN-2017-39 | en | refinedweb |
sdmx 0.2.9
Read SDMX XML files
Read SDMX XML files. I’ve only added the features I’ve needed, so this is far from being a thorough implementation. Contributions welcome.
Installation
pip install sdmx
Usage
sdmx.generic_data_message_reader(fileobj, dsd_fileobj=None, lazy=None)
Given a file-like object representing the XML of a generic data message, return a data message reader.
sdmx.compact_data_message_reader(fileobj, dsd_fileobj=None, lazy=None)
Given a file-like object representing the XML of a compact data message, return a data message reader.
Optional arguments for data message readers
- dsd_fileobj: the file-like object representing the XML of the relevant DSD. Only used if the data message does not contain a URL to the relevant DSD.
- lazy: set to True to read observations lazily to allow datasets to be read without loading the entire dataset into memory. Use with caution: lazy reading makes some assumptions about the structure of the XML (for instance, that series keys always appear before any observations in that series). These assumptions seem to be safe on files that I’ve tested, but that doesn’t mean they’re universally true.
Data message readers
Each data message reader has the following attributes:
- datasets(): returns an iterable of DatasetReader instances. Each instance corresponds to a <DataSet> element.
DatasetReader
A DatasetReader has the following attributes:
- key_family(): returns the KeyFamily for the dataset. This corresponds to the <KeyFamilyRef> element.
- series(): returns an iterable of Series instances. Each instance corresponds to a <Series> element.
KeyFamily
A KeyFamily has the following attributes:
- name(lang): the name of the key family in the language lang.
- describe_dimensions(lang): for each dimension of the key family, find the referenced concept and use its name in the language lang. Returns a list of strings in the same order as in the source file.
Series
A Series has the following attributes:
- describe_key(lang): the key of a series is a mapping from each dimension of the dataset to a value. For instance, if the dataset has a dimension named Country, the value for the series might be United Kingdom. Returns an ordered dictionary mapping strings to lists of strings. The items in the dictionary are in the same order as the dimensions returned from describe_dimensions(). For instance, if the dataset has a single dimension called Country, the returned value would be {"Country": ["United Kingdom"]}. All ancestors of a value are also described, with ancestors appearing before descendents. For instance, if the value United Kingdom has the parent value Europe, which has the parent value World, the returned value would be {"Country": ["World", "Europe", "United Kingdom"]}.
- observations(): returns an iterable of Observation instances. Each instance corresponds to an <Obs> element.
Observation
An Observation has the following attributes:
- time
- value
Example
The script below can be used to print out the values contained in a generic data message. (If you have a compact data message, then using compact_data_message_reader instead of generic_data_message_reader should also work.) Assuming the script is saved as read-sdmx-values.py, it can be used like so:
python read-sdmx-values.py path/to/generic-data-message.xml path/to/dsd.xml
import sys import sdmx def main(): dataset_path = sys.argv[1] dsd_path = sys.argv[2] with open(dataset_path) as dataset_fileobj: with open(dsd_path) as dsd_fileobj: dataset_reader = sdmx.generic_data_message_reader( fileobj=dataset_fileobj, dsd_fileobj=dsd_fileobj, ) _print_values(dataset_reader) def _print_values(dataset_reader): for dataset in dataset_reader.datasets(): key_family = dataset.key_family() name = key_family.name(lang="en") print name dimension_names = key_family.describe_dimensions(lang="en") + ["Time", "Value"] for series in dataset.series(): row_template = [] key = series.describe_key(lang="en") for key_name, key_value in key.iteritems(): row_template.append(key_value) for observation in series.observations(lang="en"): row = row_template[:] row.append(observation.time) row.append(observation.value) print zip(dimension_names, row) main()
- Author: Michael Williamson
- Keywords: sdmx
- Categories
- Development Status :: 4 - Beta
-
- Package Index Owner: michaelwilliamson
- DOAP record: sdmx-0.2.9.xml | https://pypi.python.org/pypi/sdmx/0.2.9 | CC-MAIN-2017-39 | en | refinedweb |
YAML has the ability to express hex-values, which are then decoded as numbers. However, when you want to dump a YAML document, strings will be quoted and numbers will be decimals. In order to write actual hex-values, you need to wrap your value in another type and then tell the YAML encoder how to handle it.
This is specifically possible with the ruamel YAML encoder (pypi).
An example of how to do this:
import sys import ruamel.yaml class HexInt(int): pass def representer(dumper, data): return \ ruamel.yaml.ScalarNode( 'tag:yaml.org,2002:int', '0x{:04x}'.format(data)) ruamel.yaml.add_representer(HexInt, representer) data = { 'item1': { 'string_value': 'some_string', 'hex_value': HexInt(641), } } ruamel.yaml.dump(data, sys.stdout, default_flow_style=False)
Output:
item1: hex_value: 0x0281 string_value: some_string
Please note that I require that my hex-values are two bytes and padded with zeroes, so the example above will print four characters (plus the prefix): 0x{:04x} . If this doesn’t work for you, change it to whatever you require. | https://dustinoprea.com/2018/04/ | CC-MAIN-2021-31 | en | refinedweb |
In the previous lessons, we learned about C# for loop and C# while loop. In this tutorial, we are going to learn C# do-while loop. The for loop and while checks the condition at beginning of loop then execute the statement.
The do-while loop executes the statement at least one time and checks the condition end of the loop.
In simple words, In the for loop and while loop compiler checks the condition first and execute the block of code but in the do-while loop, compiler executes the block of code once and check condition at the end of the block.
The C# while loop and C# do-while loop can say opposite each other. In the while loop, we write condition first then block of code but in the C# do-while loop, we write only do and block code. Condition writes the end of the block.
Let's understand the C# do-while loop with syntax.
do { code to be executed; } while (condition);
In the above syntax, you can see that the code of block is executed once then checks the condition. In the while loop, while condition comes before the block of code but in the do-while loop, while condition comes after the code of block end.
There is no programming without practice. Let's create an example for the do-while loop in C# programming.
In this example,we are going to printing increasing values of a variable using C# do-while loop.
using System; namespace DowhileLoop{ public class Program { public static void Main(string[] args) { int i=0; int lp=0; do{ i++; lp++; Console.WriteLine("Loop "+lp+", Value of i: "+i); } while(i<17); } } }
Output -
Loop 1, Value of i: 1 Loop 2, Value of i: 2 Loop 3, Value of i: 3 Loop 4, Value of i: 4 Loop 5, Value of i: 5 Loop 6, Value of i: 6 Loop 7, Value of i: 7 Loop 8, Value of i: 8 Loop 9, Value of i: 9 Loop 10, Value of i: 10 Loop 11, Value of i: 11 Loop 12, Value of i: 12 Loop 13, Value of i: 13 Loop 14, Value of i: 14 Loop 15, Value of i: 15 Loop 16, Value of i: 16 Loop 17, Value of i: 17
In the above C# example, we declared two variables with value(Initial value).In do-while, we first write do keyword-only then create a block of code. We make a condition at the end of do block using while keyword.
In the above example, we used ++ increment operator to increment variables value by 1 in every loop.
In the programming world, we create unique code and effective to run smoothly. In this C# example, we are going to use if-else and break statement with a do-while loop. It helps to understand making effective codes using control structures and loops in C# programming.
using System; namespace DowhileLoop{ public class Program { public static void Main(string[] args) { int a=1; do { if (a < 5) { Console.WriteLine("a is not big enough"); break; } else { Console.WriteLine("a is big enough"); break; } } while (0>a); } } }
Output -
a is not big enough
In the above example, we wrote if-else and break inside the do block and execute. In this way, you can use another control structure and loop in C# programming. | https://technosmarter.com/csharp/do-while-loop | CC-MAIN-2021-31 | en | refinedweb |
An Overview of WatchKit Tables in watchOS 2
The WatchKit Table object allows content to be displayed within a WatchKit app scene in the form of a single vertical column of rows. Tables can be used purely as a mechanism for displaying lists of information, or to implement navigation whereby the selection of a particular row within a table triggers a transition to another scene within the storyboard.
This chapter will provide an overview of tables in WatchKit, exploring how tables are structured and explaining areas such as the WKInterfaceTable class, row controllers, row controller classes, row controller types and the steps necessary to initialize a WatchKit table at runtime. The next chapter, entitled A WatchKit Table Tutorial, will then work through the creation of an example WatchKit table scene. Table based navigation will then be explored in the Implementing WatchKit Table Navigation chapter of the book.
The WatchKit Table
WatchKit tables provide a way to display information to the user in the form of a single column of rows. If a table has too many rows to fit within the watch display the user can scroll up and down within the table using the touch screen or the Digital Crown. The individual rows in a table may also be configured to call action methods when tapped by the user.
Tables are represented by the WatchKit WKInterfaceTable class, with each row displayed within the table represented by a table row controller instance.
Table Row Controller
There are two parts to the table row controller. The first is the visual representation of the row within the table. This essentially defines which user interface objects are to be displayed in the row (for example a row might consist of an Image and a Label object).
The second component of a table row is a corresponding row controller class which resides within the WatchKit app extension. This class is created as a subclass of the NSObject class and, at a minimum, contains outlets to the user interface objects contained in the row controller. Once declared, these outlets are used by the scene’s interface controller to configure the content displayed within each row. If the row controller in the scene contained two Label objects, for example, the outlets could be used to set the text on those labels for each row in the table.
Row Controller Type
The user interface objects defined within a row controller in the scene combined with the corresponding row controller class in the extension define the row controller type. A single table can consist of multiple row controller types. One row controller might, for example, contain a label and an image while another might contain two labels. The type of row controller used for each row within the table is controlled by the interface controller during the table row initialization process.
Table Row Initialization
When a scene containing a table is displayed within a WatchKit app, the table object will need a row controller instance for each row to be displayed to the user. The interface controller is responsible for performing the following initialization tasks:
1. Calculate the number of rows to be displayed in the table.
2. Request the creation of a row controller instance of a particular row controller type for each row in the table.
3. Configure the appearance of the user interface objects in each row using the outlets declared in the row controller class.
Implementing a Table in a WatchKit App Scene
A Table is added to a WatchKit scene by dragging and dropping a Table object from the Object Library onto the storyboard scene. By default, the table instance will contain a single row controller instance containing a Group object. A Group object is a single user interface element that can contain one or more interface objects in a horizontal or vertical arrangement. Figure 6-1 shows a scene with a newly added table with the default row controller:
Figure 6-1
The hierarchical structure of the table and table row controller can be viewed within the Xcode Document Outline panel. This panel appears by default to the left of the Interface Builder panel and is controlled by the small button in the bottom left hand corner (indicated by the arrow in Figure 6-2) of the Interface Builder panel.
Figure 6-2
When displayed, the document outline shows a hierarchical overview of the elements that make up a user interface layout. This enables us, for example, to see that a scene consists of a Table, Table Row Controller and a Group object:
Figure created in the table are of the same row type. This method takes as parameters the number of rows to be created and the Identifier string for the row controller type as defined in the Attributes Inspector.
- setRowTypes – Called when the table is to comprise rows of different types. This method takes as a parameter an array containing the Identifiers of the row controller types (as defined in the Attributes Inspector) in the order in which they are to appear in the table.
When the above methods are called they remove any existing rows from the table and create new rows based on the parameters provided. The methods also create an internal array containing an instance of the row controller class for each of the rows displayed in the table. These instances can then be accessed by calling the rowControllerAtIndex method of the table object. Once a reference to a row controller class instance has been obtained, the outlets declared in that instance can be used to set the attributes of the user interface objects in the row controller.
The following code listing, for example, displays a color name on each row of a table within a WatchKit app scene using the row type identified by “MyTableRowController”:
import WatchKit import Foundation class InterfaceController: WKInterfaceController { // The outlet to the Table object in the scene @IBOutlet weak var myTable: WKInterfaceTable! // The data array let colorNames = ["Red", "Green", "Blue", "Yellow"] override init() { super.init() loadTable() // Call to initialize the table rows } func loadTable() { // Create the table row controller instances // based on the number of items in colorNames array myTable.setNumberOfRows(colorNames.count, withRowType: "MyTableRowController") // Iterate through each of the table row controller // class instances. for (index, labelText) in enumerate(colorNames) { // Get a reference to the current instance let row = myTable.rowControllerAtIndex(index) as MyRowController // Set the text using the outlet in the row controller // class instance. row.myLabel.setText(labelText) } } . . }
This approach is not restricted to the initialization phase of a table. The same technique can be used to dynamically change the properties of the user interface objects in a table row at any point when a table is displayed during the lifecycle of an app. The following action method, for example, dynamically changes the text displayed on the label in row zero of the table initialized in the above code:
@IBAction func buttonTap() { let row = myTable.rowControllerAtIndex(0) as MyRowController row.myLabel.setText("Hello") }
Inserting Table Rows
Additional rows may be added to a table at runtime using the insertRowsAtIndexes method of the table instance. This method takes as parameters an index set indicating the positions at which the rows are to be inserted and the identifier of the row controller type to be used. The following code, for example, inserts new rows of type “MyImageRowController” at row index positions 0, 2 and 4:
let indexSet = NSMutableIndexSet() indexSet.addIndex(0) indexSet.addIndex(2) indexSet.addIndex(4) myTable.insertRowsAtIndexes(indexSet, withRowType: "MyImageRowController")
Removing Table Rows
Similarly, rows may be removed from a table using the removeRowsAtIndexes method of the table instance, once again passing through as a parameter an index set containing the rows to be removed. The following code, for example, removes the rows inserted in the above code fragment:
let indexSet = NSMutableIndexSet() indexSet.addIndex(0) indexSet.addIndex(2) indexSet.addIndex(4) myTable.removeRowsAtIndexes(indexSet)
Scrolling to a Specific Table Row
The table can be made to scroll to a specific row programmatically using the scrollToRowAtIndex method of the table instance, passing through as a parameter an integer value representing the index position of the destination row:
myTable.scrollToRowAtIndex(1)
A negative index value will scroll to the top of the table, while a value greater than the last index position will scroll to the end.
Summary
Tables are created in WatchKit using the WKInterfaceTable class which allows content to be presented to the user in the form of a single column of rows. Each row within a table is represented visually within a storyboard scene by a row controller which, in turn, has a corresponding row controller class residing in the app extension. A single table can display multiple row controller types so that individual rows in a table can comprise different user interface objects. Initialization and runtime configuration of the row controller instances is the responsibility of the interface controller for the scene in which the table appears. A variety of methods are available on the table class to dynamically insert and remove rows while the table is displayed to the user. | https://www.techotopia.com/index.php/An_Overview_of_WatchKit_Tables_in_watchOS_2 | CC-MAIN-2018-39 | en | refinedweb |
LDAP Directory Integration with Cisco Unity Connection
- Janice White
- 2 years ago
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Transcription
1 CHAPTER 6 LDAP Directory Integration with Cisco Unity Connection The Lightweight Directory Access Protocol (LDAP) provides applications like Cisco Unity Connection with a standard method for accessing user information that is stored in the corporate directory. Companies that centralize all user information in a single repository that is available to multiple applications can reduce maintenance costs by eliminating redundant adds, moves, and changes. Cisco Unity Connection 7.x is the first Connection release to support LDAP directory synchronization and authentication. Integrating Connection with an LDAP directory provides several benefits: User creation Connection users can be created by importing data from the LDAP directory. Data synchronization Connection can be configured to automatically synchronize user data in the Connection database with data in the LDAP directory. Single sign-on Optionally, you can configure Connection to authenticate user names and passwords for Connection web applications against the LDAP directory, so that users do not have to maintain multiple application passwords. (Phone passwords are still maintained in the Connection database.) Connection uses standard LDAPv3 for accessing data in an LDAP directory. For a list of the LDAP directories that are supported by Connection for synchronization, see the Requirements for an LDAP Directory Integration section in the System Requirements for Cisco Unity Connection Release 7.x. This chapter covers the main design issues of integrating Cisco Unity Connection 7.x with a corporate LDAP directory. See the following sections: LDAP Synchronization, page 6-1 LDAP Authentication, page 6-6 LDAP Synchronization LDAP synchronization uses an internal tool called Cisco Directory Synchronization (DirSync) to synchronize a small subset of Cisco Unity Connection user data (first name, last name, alias, phone number, and so on) with the corresponding data in the corporate LDAP directory. To synchronize user data in the Connection database with user data in the corporate LDAP directory, do the following tasks: 1. Configure LDAP synchronization, which defines the relationship between data in Connection and data in the LDAP directory. See the Configuring LDAP Synchronization section on page
2 LDAP Synchronization Chapter 6 2. Create new Connection users by importing data from the LDAP directory and/or linking data on existing Connection users with data in the LDAP directory. See the Creating Cisco Unity Connection Users section on page 6-5. For additional control, you can create an LDAP filter before you create Connection users. See the Filtering LDAP Users section on page 6-5. Configuring LDAP Synchronization Revised May 2009 When you configure LDAP directory synchronization, you can create up to five LDAP directory configurations for each Cisco Unity Connection server or cluster. Each LDAP directory configuration can support only one domain or one organizational unit (OU); if you want to import users from five domains or OUs, you must create five LDAP directory configurations. A Connection Digital Network also supports up to five LDAP directory configurations for each Connection server or cluster joined to the network. For example, if you have a Digital Network with five servers, you can import users from up to 25 domains. In each LDAP directory configuration, you specify: The user search base that the configuration will access. A user search base is the position in the LDAP directory tree where Connection begins its search for user accounts. Connection imports all users in the tree or subtree (domain or OU) specified by the search base. A Connection server or cluster can only import LDAP data from subtrees with the same directory root, for example, from the same Active Directory forest. If you are using an LDAP directory other than Microsoft Active Directory, and if you create a Connection LDAP directory configuration that specifies the root of the directory as the user search base, Connection will import data for every user in the directory. If the root of the directory contains subtrees that you do not want Connection to access (for example, a subtree for service accounts), you should do one of the following: Create two or more Connection LDAP directory configurations, and specify search bases that omit the users that you do not want Connection to access. Create an LDAP search filter. For more information, see the Filtering LDAP Users section in the Integrating Cisco Unity Connection with an LDAP Directory chapter of the System Administration Guide for Cisco Unity Connection Release 7.x. For directories other than Active Directory, we recommend that you specify user search bases that include the smallest possible number of users to speed synchronization, even when that means creating multiple configurations. If you are using Active Directory and if a domain has child domains, you must create a separate configuration to access each child domain; Connection does not follow Active Directory referrals during synchronization. The same is true for an Active Directory forest that contains multiple trees you must create at least one configuration to access each tree. In this configuration, you must map the UserPrincipalName (UPN) attribute to the Connection Alias field; the UPN is guaranteed by Active Directory to be unique across the forest. For additional considerations on the use of the UPN attribute in a multi-tree AD scenario, see the Additional Considerations for Authentication and Microsoft Active Directory section on page 6-8. If you are using Digital Networking to network two or more Connection servers that are each integrated with an LDAP directory, do not specify a user search base on one Connection server that overlaps a user search base on another Connection server, or you will have user accounts and mailboxes for the same Connection user on more than one Connection server. 6-2
3 Chapter 6 LDAP Synchronization Note You can eliminate the potential for duplicate users by creating an LDAP filter on one or more Connection servers. See the Filtering LDAP Users section in the Integrating Cisco Unity Connection with an LDAP Directory chapter of the System Administration Guide for Cisco Unity Connection Release 7.x. The administrator account in the LDAP directory that Connection will use to access the subtree specified in the user search base. Connection performs a bind to the directory and authenticates by using this account. We recommend that you use an account dedicated to Connection, with minimum permissions set to read all user objects in the search base and with a password set never to expire. (If the password for the administrator account changes, Connection must be reconfigured with the new password.) If you create more than one configuration, we recommend that you create one administrator account for each configuration and give that account permission to read all user objects only within the corresponding subtree. When creating the configuration, you enter the full distinguished name for the administrator account; therefore the account can reside anywhere in the LDAP directory tree. The frequency with which Connection automatically resynchronizes the Connection database with the LDAP directory, if at all. You can specify the date and time of the next resynchronization, whether the resynchronization occurs just once or on a schedule and, if on a schedule, what you want the frequency to be in hours, days, weeks, or months (with a minimum value of six hours). We recommend that you stagger synchronization schedules so that multiple agreements are not querying the same LDAP servers simultaneously. Schedule synchronization to occur during nonbusiness hours. The port on the LDAP server that Connection uses to access LDAP data. Optionally, whether to use SSL to encrypt data that is transmitted between the LDAP server and the Connection server. One or more LDAP servers. For some LDAP directories, you can specify up to three LDAP directory servers that Connection uses when attempting to synchronize. Connection tries to contact the servers in the order that you specify. If none of the directory servers responds, synchronization fails; Connection tries again at the next scheduled synchronization time. You can use IP addresses rather than host names to eliminate dependencies on Domain Name System (DNS) availability. Note Not all LDAP directories support specifying additional LDAP directory servers to act as backup in case the LDAP directory server that Connection accesses for synchronization becomes unavailable. For information on whether your LDAP directory supports specifying multiple directory servers, see the Requirements for an LDAP Directory Integration section in the System Requirements for Cisco Unity Connection Release 7.x. The mapping of LDAP directory attributes to Connection fields, as listed in Table 6-1. Note that the mapping to the Connection Alias field must be the same for all configurations. As you choose an LDAP attribute to map to the Connection Alias field: Confirm that every user that you want to import from the LDAP directory into Connection has a unique value for that attribute. If there are already users in the Connection database, confirm that none of the users that you want to import from the directory has a value in that attribute that matches the value in the Alias field for an existing Connection user. 6-3
4 LDAP Synchronization Chapter 6 Note that for every user that you want to import from the LDAP directory into Connection, the LDAP sn attribute must have a value. Any LDAP user for whom the value of the sn attribute is blank will not be imported into the Connection database. To protect the integrity of data in the LDAP directory, you cannot use Connection tools to change any of the values that you import. Connection-specific user data (for example, greetings, notification devices, conversation preferences) is managed by Connection and stored only in the local Connection database. Note that no passwords or PINs are copied from the LDAP directory to the Connection database. If you want Connection users to authenticate against the LDAP directory, see the LDAP Authentication section on page 6-6. Table 6-1 Mapping of LDAP Directory Attributes to Cisco Unity Connection User Fields LDAP Directory Attribute One of the following: samaccountname mail employeenumber telephonenumber userprinciplename givenname One of the following: middlename initials SN manager department One of the following: telephonenumber ipphone One of the following: mail samaccountname title homephone mobile pager Cisco Unity Connection User Field Alias First Name Initials Last Name Manager Department Corporate Phone Corporate Address Title Home (imported but not currently used, and not visible in Connection Administration) Mobile (imported but not currently used, and not visible in Connection Administration) Pager (imported but not currently used, and not visible in Connection Administration) 6-4
5 Chapter 6 LDAP Synchronization When clustering (active/active high availability) is configured, all user data, including data imported from the LDAP directory, is automatically replicated from the Connection publisher server to the subscriber server. In this configuration, the Cisco DirSync service runs only on the publisher server. Creating Cisco Unity Connection Users On a Cisco Unity Connection system that is integrated with an LDAP directory, you can create Connection users by importing data from the LDAP directory, converting existing Connection users to synchronize with the LDAP directory, or both. Note the following: When you create Connection users by importing LDAP data, Connection takes the values specified in Table 6-1 from the LDAP directory and fills in the remaining information from the Connection user template that you specify. When you convert existing users, existing values in the fields in Table 6-1 are replaced with the values in the LDAP directory. For any user that you want to import from the LDAP directory, the value in the LDAP attribute that maps to the Connection Alias field cannot match the value in the Connection Alias field for any Connection object (standalone users, users already imported from an LDAP directory, users imported from Cisco Unified Communications Manager via AXL, contacts, distribution lists, and so on). After you have synchronized Connection with the LDAP directory, you can continue to add Connection users who are not integrated with the LDAP directory. You can also continue to add Connection users by importing users from Cisco Unified Communications Manager via an AXL Server. After you have synchronized Connection with the LDAP directory, new LDAP directory users are not automatically imported into Connection, but must be imported manually. After a user has been imported from LDAP, the user page in Cisco Unity Connection Administration identifies the user as an Active User Imported from LDAP Directory. Subsequently when changes are made to user data in the corporate directory, Connection fields that are populated from the LDAP directory are updated with the new LDAP values during the next scheduled resynchronization. Filtering LDAP Users Revised May 2009 You may want additional control over which LDAP users you import into Cisco Unity Connection for a variety of reasons. For example: The LDAP directory has a flat structure that you cannot control sufficiently by specifying user search bases. You only want a subset of LDAP user accounts to become Connection users. The LDAP directory structure does not match the way you want to import users into Connection. For example: If organizational units are set up according to an organizational hierarchy but users are mapped to Connection by geographical location, there might be little overlap between the two. If all users in the directory are in one tree or domain but you want to install more than one Connection server, you need to do something to prevent users from having mailboxes on more than one Connection server. 6-5
6 LDAP Authentication Chapter 6 In these cases, you may want to use the set cuc ldapfilter CLI command to provide additional control over user search bases. Note the following: The set cuc ldapfilter CLI command cannot be used with Cisco Unified CMBE. You can only create one filter per Connection server or Connection cluster pair, so the LDAP filter must specify all of the users that you want to synchronize with Connection users. When you configure LDAP synchronization in Connection, you can further filter the LDAP users by your choice of user search bases. The filter must adhere to the LDAP filter syntax specified in RFC 2254, The String Representation of LDAP Search Filters. The filter syntax is not verified, and no error message is returned. We recommend that you verify the LDAP filter syntax before you include it in this command. If you re-run this command and specify a filter that excludes some of the users who were accessible with the previous filter, the Connection users who are associated with the now-inaccessible LDAP users will be converted to standalone Connection users over the next two scheduled synchronizations or within 24 hours, whichever is greater. The users will still be able to log on to Connection by using the telephone user interface, callers can still leave messages for them, and their messages will not be deleted. However, they will not be able to log on to Connection web applications while Connection is converting them to standalone users, and after they have become standalone users, their web-application passwords will be the passwords that were assigned when their Connection accounts were created. LDAP Authentication Some companies want the convenience of single sign-on credentials for their applications. To authenticate logons to Connection web applications against user credentials in an LDAP directory, you must synchronize Connection user data with user data in the LDAP directory as described in the LDAP Synchronization section on page 6-1. Only passwords for Connection web applications (Cisco Unity Connection Administration for administration, Cisco Personal Communications Assistant for end users), and for IMAP applications that are used to access Connection voice messages, are authenticated against the corporate directory. You manage these passwords by using the administration application for the LDAP directory. When authentication is enabled, the password field is no longer displayed in Cisco Unity Connection Administration.) For telephone user interface or voice user interface access to Connection voice messages, numeric passwords (PINs) are still authenticated against the Connection database. You manage these passwords in Connection Administration, or users manage them in the Cisco PCA. The LDAP directories that are supported for LDAP authentication are the same as those supported for synchronization. See the Requirements for an LDAP Directory Integration section in the System Requirements for Cisco Unity Connection Release 7.x. See the following sections for additional details: Configuring LDAP Authentication, page 6-7 How LDAP Authentication Works, page 6-7 Additional Considerations for Authentication and Microsoft Active Directory, page
7 Chapter 6 LDAP Authentication Configuring LDAP Authentication Configuring LDAP authentication is much simpler than configuring synchronization. You specify only the following: A user search base. If you created more than one LDAP configuration, when you configure authentication, you must specify a user search base that contains all of the user search bases that you specified in your LDAP configurations. The administrator account in the LDAP directory that Cisco Unity Connection will use to access the search base. We recommend that you use an account dedicated to Connection, with minimum permissions set to read all user objects in the search base and with a password set never to expire. (If the password for the administrator account changes, Connection must be reconfigured with the new password.) You enter the full distinguished name for the administrator account; therefore the account can reside anywhere in the LDAP directory tree. One or more LDAP servers. You can specify up to three LDAP directory servers that Connection uses when attempting to authenticate. Connection tries to contact the servers in the order that you specify. If none of the directory servers responds, authentication fails. You can use IP addresses rather than host names to eliminate dependencies on Domain Name System (DNS) availability. How LDAP Authentication Works When LDAP synchronization and authentication are configured in Cisco Unity Connection, authenticating the alias and password of a user against the corporate LDAP directory works as follows: 1. A user connects to the Cisco Personal Communications Assistant (PCA) via HTTPS and attempts to authenticate with an alias (for example, jsmith) and password. 2. Connection issues an LDAP query for the alias jsmith. For the scope for the query, Connection uses the LDAP search base that you specified when you configured LDAP synchronization in Cisco Unity Connection Administration. If you chose the SSL option, the information that is transmitted to the LDAP server is encrypted. 3. The corporate directory server replies with the full Distinguished Name (DN) of user jsmith, for example, cn=jsmith, ou=users, dc=vse, dc=lab. 4. Connection attempts an LDAP bind by using this full DN and the password provided by the user. 5. If the LDAP bind is successful, Connection allows the user to proceed to the Cisco PCA. If all of the LDAP servers that are identified in a Connection LDAP directory configuration are unavailable, authentication for Connection web applications fails, and users are not allowed to access the applications. However, authentication for the telephone and voice user interfaces will continue to work, because these PINs are authenticated against the Connection database. When the LDAP user account for a Connection user is disabled or deleted, or if an LDAP directory configuration is deleted from the Connection system, the following occurs: 1. Initially, when Connection users try to log on to a Connection web application, LDAP authentication fails because Connection is still trying to authenticate against the LDAP directory. If you have multiple LDAP directory configurations accessing multiple LDAP user search bases, and if only one configuration was deleted, only the users in the associated user search base are affected. Users in other user search bases are still able to log on to Connection web applications. 2. At the first scheduled synchronization, users are marked as LDAP inactive in Connection. Attempts to log on to Connection web applications continue to fail. 6-7
8 LDAP Authentication Chapter 6 3. At the next scheduled synchronization that occurs at least 24 hours after users are marked as LDAP inactive, all Connection users whose accounts were associated with LDAP accounts are converted to Connection standalone users. For each Connection user, the password for Connection web applications and for IMAP access to Connection voice messages becomes the password that was stored in the Connection database when the user account was created. (This is usually the password in the user template that was used to create the user.) Connection users do not know this password, so an administrator must reset it. The numeric password (PIN) for the telephone user interface and the voice user interface remains unchanged. Note the following regarding Connection users whose LDAP user accounts were disabled or deleted, or who were synchronized via an LDAP directory configuration that was deleted from Connection: The users can continue to log on to Connection by phone during the period in which Connection is converting them from an LDAP-synchronized user to a standalone user. Their messages are not deleted. Callers can continue to leave messages for these Connection users. Additional Considerations for Authentication and Microsoft Active Directory When you enable LDAP authentication with Active Directory, we recommend that you configure Cisco Unity Connection to query an Active Directory global catalog server for faster response times. To enable queries against a global catalog server, in Connection Administration, specify the IP address or host name of a global catalog server. For the LDAP port, specify either 3268 if you are not using SSL to encrypt data that is transmitted between the LDAP server and the Connection server, or 3269 if you are using SSL. Using a global catalog server for authentication is even more efficient if the users that are synchronized from Active Directory belong to multiple domains, because Connection can authenticate users immediately without having to follow referrals. For these cases, configure Connection to access a global catalog server, and set the LDAP user search base to the top of the root domain. A single LDAP user search base cannot include multiple namespaces, so when an Active Directory forest includes multiple trees, Connection must use a different mechanism to authenticate users. In this configuration, you must map the LDAP userprincipalname (UPN) attribute to the Connection Alias field. Values in the UPN attribute, which look like addresses must be unique in the forest. Note When an Active Directory forest contains multiple trees, the UPN suffix (the part of the address after symbol) for each user must correspond to the root domain of the tree where the user resides. If the UPN suffix does not match the namespace of the tree, Connection users cannot authenticate against the entire Active Directory forest. However, you can map a different LDAP attribute to the Connection Alias field and limit the LDAP integration to a single tree within the forest. For example, suppose an Active Directory forest contains two trees, avvid.info and vse.lab. Suppose also that each tree includes a user whose samaccountname is jdoe. Connection authenticates a logon attempt for jdoe in the avvid.info tree as follows: 1. The user jdoe connects to the Cisco Personal Communications Assistant (PCA) via HTTPS and enters a UPN and password. 6-8
9 Chapter 6 LDAP Authentication 2. Connection performs an LDAP query against an Active Directory global catalog server by using the UPN. The LDAP search base is derived from the UPN suffix. In this case, the alias is jdoe and the LDAP search base is dc=avvid, dc=info. 3. Active Directory finds the Distinguished Name corresponding to the alias in the tree that is specified by the LDAP query, in this case, cn=jdoe, ou=users, dc=avvid, dc=info. 4. Active Directory responds via LDAP to Connection with the full Distinguished Name for this user. 5. Connection attempts an LDAP bind by using the Distinguished Name and the password initially entered by the user. 6. If the LDAP bind is successful, Connection allows the user to proceed to the Cisco PCA. 6-9
10 LDAP Authentication Chapter
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Content Filtering Client Policy & Reporting Administrator s Guide
Content Filtering Client Policy & Reporting Administrator s Guide Notes, Cautions, and Warnings NOTE: A NOTE indicates important information that helps you make better use of your system. CAUTION: A CAUTION | http://docplayer.net/14071686-Ldap-directory-integration-with-cisco-unity-connection.html | CC-MAIN-2018-39 | en | refinedweb |
Good all day! Error-D:\Classes\mutex1\mutex1.cpp | 9 | error: ' mutex ' in namespace ' std ' does not name a type |
Something is wrong with my compiler. So far, everything has always been well compiled, worked remarkably well.
Various advice and research problems I accept with joy.
I act through CodeBlocks and my computer has 32 bits.
Downloaded D: \ gcc-6.3.0 \ gcc-6.3.0 (gcc-6.3.0.tar.bz2)
It can somehow be connected in the right places?
Is not it necessary to compile something at first?
In general, I imagine the steps of correcting the problem rather poorly. What additional libraries are available?
Last edited by Дмитро; August 5th, 2017 at 04:01 PM.
Did you try to search for this problem with google?
Victor Nijegorodov
Originally Posted by VictorN
Did you try to search for this problem with google?
Loool Google is a man's best friend
Forum Rules | http://forums.codeguru.com/showthread.php?559983-mutex-in-namespace-std-does-not-name-a-type&s=4f889e0b06c9c63633cbaf56ee3a59c9&p=2216641 | CC-MAIN-2018-39 | en | refinedweb |
Android Dice Roller Source Code for Apps
An Android Dice Roller tutorial with source code for a basic dice roller app. OK it might not be a fancy 3D dice roller but it'll get you going. You can always upgrade your dice roller code later. Simply add the example source code to any app that requires a dice roll. Dice face images and a 3D dice icon image are provided, all the images and code are free to use and royalty free. The code here is for a six sided dice but can be adapted for larger dice or more than one dice. There are no restrictions placed upon the code or the dice face images (all the images are free as they are in the public domain).
By the way (BTW) I know that dice is the plural and means two or more die, but it is common to refer to one die as a dice so we will be doing that here.
(This Android dice roller tutorial assumes that Android Studio is installed, a basic App can be created and run, and the code in this article can be correctly copied into Android Studio. The example code can be changed to meet your own requirements. When entering code in Studio add import statements when prompted by pressing Alt-Enter.)
The Dice Face Images and Sound
The code given in this article is for a common six sided dice. A dice is a cube, and each side of the cube (each face) has one of the numbers from one to six. Here six PNG images are used to show the dice roll result. Plus there is a 3D image shown for the dice roll. The same 3D image is used for the app icon. The dice images are from Open Clip Art Library user rg1024.
The sound of a dice roll is stored in shake_dice.mp3. It is by Mike Koenig and is from SoundBilble.com.
All the resources ready to add into a Android Studio project are available from this dice resources Zip file.
Android Studio Dice Roller Project
For this tutorial start a new Android Studio project. Here the Application name is Dice and an Empty Activity is used with other settings left at default values. Add the dice resources (from above) to the project by copying them to the res folder. Studio will update the Project explorer automatically (or use the Synchronize option).
Optionally change the app icon entries in the AndroidManifest.xml file to dice3d and dice3d_rounded.png, i.e.
android:icon="@drawable/dice3d" and
android:roundIcon="@mipmap/dice3d_rounded".
Android Dice Roller Project Layout
The screen for the dice roller is very simple. Just an
ImageView which when pressed the roll is performed. If using the basic starting empty screen then open the activity_main.xml layout file. Delete the Hello World! default
TextView (if present). From the widget Pallete, under Images, drag and drop an ImageView onto the layout and select dice3droll on the Resources chooser dialog. (Add the constraints if using a
ConstraintLayout.)
The layout source should be similar to the following:
<?xml version="1.0" encoding="utf-8"?> <android.support.constraint.ConstraintLayout xmlns: <ImageView android: </android.support.constraint.ConstraintLayout>
Android Dice Roller Source Code
A Java
Random class (to generate the dice numbers) is declared along with an Android
SoundPool (to play the dice roll sound). To load the right picture for each roll a switch statement is used. Some feedback is provided to the user so that they can see that a new roll has occurred. An identical number can be rolled in succession. To provide the feedback a 3D dice image will be displayed. However, because the roll happens so fast (unlike a real dice) the 3D image would not be seen. Therefore a
Timer is used to provide a delay. This allows the UI to update. The Timer sends a
Handler message signal to a
Callback to perform the roll (the same method as described in the post Why Your TextView or Button Text Is Not Updating). Finally the roll value is used to update the UI.
Update: The creation of the SoundPool object has changed due to the deprecation of the constructor in Android API 21 and later. The code has been changed to allow for this, using a
SoundPool.Builder. To support pre-Lollipop Android versions a new class is added to the project called PreLollipopSoundPool. The code for this new class follows this MainActivity class:
package com.example.dice; import android.media.AudioAttributes; import android.media.SoundPool; import android.os.Build; import android.os.Handler; import android.os.Message; import android.support.v7.app.AppCompatActivity; import android.os.Bundle; import android.view.View; import android.widget.ImageView; import java.util.Random; import java.util.Timer; import java.util.TimerTask; public class MainActivity extends AppCompatActivity { ImageView dice_picture; //reference to dice picture Random rng=new Random(); //generate random numbers SoundPool dice_sound; //For dice sound playing int sound_id; //Used to control sound stream return by SoundPool Handler handler; //Post message to start roll Timer timer=new Timer(); //Used to implement feedback to user boolean rolling=false; //Is die rolling? @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_main); //Our function to initialise sound playing InitSound(); //Get a reference to image widget dice_picture = (ImageView) findViewById(R.id.imageView); dice_picture.setOnClickListener(new HandleClick()); //link handler to callback handler=new Handler(callback); } //User pressed dice, lets start private class HandleClick implements View.OnClickListener { public void onClick(View arg0) { if (!rolling) { rolling = true; //Show rolling image dice_picture.setImageResource(R.drawable.dice3droll); //Start rolling sound dice_sound.play(sound_id, 1.0f, 1.0f, 0, 0, 1.0f); //Pause to allow image to update timer.schedule(new Roll(), 400); } } } //New code to initialise sound playback void InitSound() { if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.LOLLIPOP) { //Use the newer SoundPool.Builder //Set the audio attributes, SONIFICATION is for interaction events //uses builder pattern AudioAttributes aa = new AudioAttributes.Builder() .setUsage(AudioAttributes.USAGE_ASSISTANCE_SONIFICATION) .setContentType(AudioAttributes.CONTENT_TYPE_SONIFICATION) .build(); //default max streams is 1 //also uses builder pattern dice_sound= new SoundPool.Builder().setAudioAttributes(aa).build(); } else { //Running on device earlier than Lollipop //Use the older SoundPool constructor dice_sound=PreLollipopSoundPool.NewSoundPool(); } //Load the dice sound sound_id=dice_sound.load(this,R.raw.shake_dice,1); } //When pause completed message sent to callback class Roll extends TimerTask { public void run() { handler.sendEmptyMessage(0); } } //Receives message from timer to start dice roll Handler.Callback callback = new Handler.Callback() { public boolean handleMessage(Message msg) { //Get roll result //Remember nextInt returns 0 to 5 for argument of 6 //hence + 1 switch(rng.nextInt(6)+1) { case 1: dice_picture.setImageResource(R.drawable.one); break; case 2: dice_picture.setImageResource(R.drawable.two); break; case 3: dice_picture.setImageResource(R.drawable.three); break; case 4: dice_picture.setImageResource(R.drawable.four); break; case 5: dice_picture.setImageResource(R.drawable.five); break; case 6: dice_picture.setImageResource(R.drawable.six); break; default: } rolling=false; //user can press again return true; } }; //Clean up protected void onPause() { super.onPause(); dice_sound.pause(sound_id); } protected void onDestroy() { super.onDestroy(); timer.cancel(); } }
To support APIs pre Lollipop add a new class called PreLollipopSoundPool in the same folder as the MainActivity class. Creating it as a separate class supports Java lazy loading, ensuring the app works when the deprecated method is removed from Android.
package com.example.dice; import android.media.AudioManager; import android.media.SoundPool; /** * Created a pre Lollipop SoundPool */ final class PreLollipopSoundPool { @SuppressWarnings("deprecation") public static SoundPool NewSoundPool() { return new SoundPool(1, AudioManager.STREAM_MUSIC,0); } }
The Android dice roller source code is now ready to run, either in an AVD or an actual Android device.
The complete source code is available in android_dice_roller.zip or from the Android Example Projects page. The code can be seen running using Dice.apk, also available from the projects page.
See Also
- The dice images were from the Open Clip Art Library by rg1024.
- The dice roll sound is from the SoundBible.com.
- See the other Android Studio example projects to learn Android app programming.
Archived Comments
Phil said on 27/09/2014: Nice code. I’m hoping to write a more complex Roller, and this is very useful!
Cesar said on 18/07/2013: I don't get it. I have tried and tried, but I don't get it. Line
handler=new Handler(callback); shows me an error. callback execution doesn't enter
handleMessage(Message msg). If I initialize
new Handler(callback); in other ways the app crashes. Can you help me, please? Thanks a lot!
Tek Eye said, in reply to Cesar, on 18/07/2013: Hi, the Handler class used here is from android.os and not java.util.logging, Callback is from android.os as well, though you probably used the correct imports. To help I’ve added a zip file with all the source code to the Android Example Projects page as well as a link in this article. You can import the code from the Zip file into a new Android project and run it to see it working. Failing that there is an APK file on the same page to install onto a device to see the code working (link to APK in article as well). Note you may need to change your device settings temporarily to allow the APK to be installed, only install APKs from trusted sources.
Author:Daniel S. Fowler Published: Updated: | https://tekeye.uk/android/examples/android-dice-code | CC-MAIN-2018-39 | en | refinedweb |
SBT
Importing an sbt project
- Click Import Project or Open on the welcome screen.
- In the dialog that opens, select a directory that contains your sbt project or simply
build.sbt. Click OK.
- Follow the steps suggested in the Import Project wizard.
You can use the suggested default settings since they are enough to successfully import your project.
We recommend that you enable the Use sbt shell for build and import (requires sbt 0.13.5+) option when you use code generation or other features that modify the build process in sbt. If your sbt project is not in the IntelliJ IDEA project root directory, we suggest you skip this option.
You can also select the appropriate option for grouping modules in your project.
Ensuring sbt and Scala versions compatibility
Often you share your project across a team and need to use a specific version of sbt.
You can override the sbt version in your project's
build.properties file.
- Create or import your sbt project.
- In the Project tool window, in the source root directory locate the
build.propertiesfile and open it in the editor.
- In the editor explicitly specify the version of sbt that you want to use in the project.
sbt.version=xxx
- Refresh your project. (Click the
in the sbt projects tool window.)
Managing sbt projects
sbt project structure
When you create or import an sbt project, IntelliJ IDEA generates the following sbt structure:
- sbt project (proper build) which defines a project and contains
build.sbtfile, src, and target directories, modules; anything related to a regular project.
- sbt build project which is defined in the project subdirectory. It contains additional code that is part of the build definition.
- The sbt projects tool window which contains sbt tasks, commands, and settings that you can execute.
When you work with sbt projects you use the
build.sbt file to make main changes to your project since IntelliJ IDEA considers an sbt configuration as a single source of truth.
Adding a library to the sbt project
You can add sbt dependencies via the
build.sbt file or you can use the
import statement in your
.scala file.
- Open a
.scalafile in the editor.
- Specify a library you want to import.
- Put the cursor on the unresolved package and press Alt+Enter.
- From the list of available intention actions, select Add sbt dependency.
- Follow the steps suggested in the wizard that opens and click Finish.
- IntelliJ IDEA downloads the artifact, adds the dependency to the
build.sbtfile and to the sbt projects tool window.
- As soon as IntelliJ IDEA detects changes in
build.sbt, a notification suggesting to refresh your project will appear. Refresh your project. (Click the
in the sbt projects tool window.)
Alternatively, use the auto-import option located in the sbt settings.
Working with sbt shell
An sbt shell is embedded in the sbt project and is available on your project start. You can use the sbt shell for executing sbt commands and tasks, for running, and debugging your projects.
- To start the sbt shell, press Ctrl+Shift+S(for Windows) or Cmd+Shift+S (for Mac OS X). Alternatively, click
on the toolbar located on the bottom of the screen.
- To use the sbt shell for build and import procedures, select the Use sbt shell for build and import (requires sbt 0.13.5+) option located in the sbt settings and perform steps described in the Run a Scala application using the sbt shell section. Note that sbt versions 0.13.16.+ / 1.0.3.+ are recommended.
- To use the sbt shell for debugging, refer to the debugging with sbt shell section.
- To run your tests from the sbt shell:
- Open a run/debug configuration ().
- Create a test configuration and select the use sbt option from the available settings.
Running sbt tasks
- You can run sbt tasks by selecting the one you need from the the sbt Tasks directory in the sbt projects tool window.
- You can manually enter your task (code completion is supported) in the sbt shell and run it directly from there.
- You can create a run configuration for a task. For example, you can create a custom task which is not part of the list of tasks located in the sbt projects tool window.
- Open (Shift+Alt+F10) a run configuration.
- Specify the run configuration settings and click OK. If you need, you can add another configuration or a task to execute before running your configuration. Click the
icon in the Before Launch section and from the list that opens select what you need to execute.
IntelliJ IDEA displays results in the sbt shell window.
Working with sbt settings
To access sbt settings, click
in the sbt projects tool window.
You can use sbt settings for the following notable actions:
- If you want sbt automatically refresh your project every time you make changes to
build.sbt, select Use auto-import.
- To delegate running builds to sbt, select Use sbt shell for build and import.
- To debug your code via the sbt shell, select Enable debugging for sbt shell option that enables a debug button (
) in the sbt shell tool window. To start the debugging session, simply click this button. For more information on debugging, see debugging with sbt.
- To change the .ivy cache location in your project or set other sbt properties, use the VM parameters field.
To check the most common sbt issues and workarounds, see the sbt troubleshooting section. | https://www.jetbrains.com/help/idea/2018.2/sbt-support.html | CC-MAIN-2018-39 | en | refinedweb |
0
I'm learning C++ from the book C++ Primer Plus. At the end of the chapter on loops, they want us to design a structure, allocate adequate memory for an array of such structures using new, then feed input data to it. I got this code:
#include <iostream> using namespace std; struct car{ char make[20]; int yearMade; }; int main() { int count = 0; cout<<"How many cars do you wish to catalog today? "; cin>>count; car * autocar = new car[count]; for(int i=1; i<=count; i++){ cout<<"Car #"<<count<<": "<<endl; cout<<"Please enter the make: "; cin.get(autocar[i-1]->make).get(); cout<<"Please enter the year manufactured: "; cin>>autocar[i-1]->yearMade; cout<<endl; } cout<<endl<<"Here is your collection: "<<endl; for(int i = 1; i<=count; i++){ cout<<autocar[i-1]->yearMade<<" "<<autocar[i-1]->make<<endl; } delete [] autocar; return 0; }
Now, the error comes on line 21, 24 and 29. The compiler is complaining that the "base operand of '->' has non-pointer type 'car'".
Beats me. I made sure that the type is a pointer, as you can see in the statement in line 16. Any suggestions? | https://www.daniweb.com/programming/software-development/threads/306841/a-problem-with-the-membership-operator | CC-MAIN-2018-39 | en | refinedweb |
The previous part of the series briefly explained how to implement a Custom Control library for SAPUI5 and then deploy it to the HCP to make it possible using the library by your other UI5 apps inside the Web IDE.
Deployment to on-premise SAP ABAP Repository
In Part 1 I unequivocally stated that all the library files should be put into the project root folder. However Web IDE works equally fine with the folder hierarchy according to the library name space. So, when first trying to load deployed library within Web IDE I created the name space hierarchy starting with src folder and thus the new-app.json file (or more precisely the entry in “routes” array) looked like this:
{ "path": "/webapp/resources/", "target": { "type": "application", "name": "mycustomlib", "entryPath": "/src/my/custom/control" }, "description": "Custom control library" }
But the next question is: how we put the library onto SAP ABAP server?
As you might know SAP by default stores all the SAPUI5 library files in the MIME repository accessible from SE80. If you select uppermost long button “MIME repository” and then select in the tree-pane this folder path “/SAP/PUBLIC/BC/UI5/LIBRARIES/VER” you should see something like this:
And under each version folder (e.g. “1.38”) you can find all the resources of SAPUI5 library.
My first (quite naive) plan was to mimic the similar structure and place my custom library to the same MIME repository along with other SAPUI5 resources. All the tries failed with error 404.
And then I decided to try the similarity principle: if the Web IDE managed to load library from deployed HCP application, then probably SAP NW server should do the same with library deployed as BSP application. And so it goes. The library with accurately placed folder hierarchy (see above) was deployed to the SAP ABAP Repository under the name “zcustlib“… and again error 404!
However looking into the network log in Chrome development tools I have noticed that the faulty request url which tried to load library control file “my.custom.control.ProductRating.js” looks like this: http://…/sap/bc/ui5_ui5/sap/zcustlib/~C2BA8580F0AD68211B5F48EF79A345B7~5/ProductRating.js.
Just notice that SAP NW server successfully mapped the library name space “my.custom.control” into the deployed application name “zcustlib“. Hope? Yes.
It appears that for one or another reason SAP NW UI5 runtime just ignores the library namespace folder hierarchy and always tries to load library files from the root folder of the deployed BSP application. Probably there is some other configuration file to overcome this, for example UI5RepositoryPathMapping.xml, but I could not find a proper documentation on this and I left it as it is.
Next time I put all the library files into the project root folder and remapped the corresponding entry in the neo-app.json file as described in Part 1. After that all apps started working both in Web IDE and in SAP ABAP Repository.
Theming and CSS
If your library has its own CSS styles then according to the guide you have to create library.css file for each supported theme. To do this you create themes folder within the folder where your library js files reside. Then inside themes folder you create subfolder for each supported theme and put there the corresponding library.css file and also “base” folder which should contain default styles, as on the following figure:
This should work in theory, while in practice the Web IDE managed to load control js files but made no attempt to load library.css – there were no trace of such action in the Network log.
After another set of try-and-fail attempts I finally figured out how to make Web IDE load the library CSS file: add an entry into dependencies in the manifest.json file of the app project. That’s the place where you normally enter other library dependencies such as sap.ui.core or sap.m. For our custom library the entry should look like this:
"sap.ui5": { ... "dependencies": { "minUI5Version": "1.36.0", "libs": { "sap.ui.core": { "minVersion": "1.36.0" }, "sap.m": { "minVersion": "1.36.0" }, "my.custom.control": { } } }, ...
With this setup the app runs fine from the Web IDE, however after I attempted to deploy it to SAP ABAP Repository it refused to start with 404 error. It appears that if you put the library into the manifest.json dependency array SAP NW server always tries to load library-preload.js file from library root folder – this is a minified file containing all the library modules. Unlike an ordinary UI5 app project (where Web IDE creates Component-preload.js file) it does not do this for library project.
Another round of source code investigation and debugging finally produced a workaround: to add library.js file as a dependency to each of library controls. After a minor dependency part amendment our ProductRating.js file header should look like this:
sap.ui.define([ "sap/ui/core/Control", "sap/m/RatingIndicator", "sap/m/Label", "sap/m/Button", "./library" // Load library.js as a dependency ], function (Control, RatingIndicator, Label, Button) { ... });
As you can see we don’t need a parameter for the library.js dependency because function sap.ui.define already does what we need: it executes the loaded module to obtain an object instance. In case of library.js module one thing it has to do is calling Core function initLibrary, and otherwise it is just a common executable JavaScript module. Thus, we can add a call to Core function includeLibraryTheme and the resulting library.js file would look as follows:
/* global my:true */ sap.ui.define([ "jquery.sap.global", "sap/ui/core/library" ], // library dependency function(jQuery) { "use strict"; sap.ui.getCore().includeLibraryTheme("my.custom.control"); sap.ui.getCore().initLibrary({ name: "my.custom.control", version: "1.0.0", dependencies: ["sap.ui.core"], types: [], interfaces: [], controls: [ "my.custom.control.ProductRating" ], elements: [] }); return my.custom.control; }, /* bExport= */ false);
Bottom line
So, let’s summarise all the findings to produce a workable setup for shareable custom control library:
- Create a separate project for the library in the Web IDE and put all the library files into the project root folder
- Make sure you marked the type of the project as “library” not “application” in the manifest.json file (see Part 1)
- Fill the “id” value of “sap.app” part of the manifest.json file with the library name space (see Part 1)
- If the library has CSS rules create proper folder hierarchy starting with themes folder and place the library.css file into each supported theme folder
- If the library has CSS rules add includeLibraryTheme Core function call into the library.js file
- Add the library to the dependency list of each library control files
- Deploy the library project both to HCP and SAP ABAP Repository
- In the app project which uses the library amend the neo-app.json file as suggested in the Part 1
- Don’t add the library to the manifest.json dependency of the app, otherwise SAP NW server can refuse to load the library
- Have fun
Further steps
As you see this is quite a minimal setup allowing you to use your own custom control library both within Web IDE and on SAP ABAP Repository. At the same time there are still tasks to make it a kind of state-of-the-art solution:
- It is unclear if it is possible at all to make a custom library shareable for more than one HCP account (just like SAPUI5)
- At this point Web IDE does not “build” the library project – it would be good to have a minified version of the library files together with library-preload.js file
- As far as I can understand UI Theme designer currently does not support generation of library specific css files. UI Theme designer produces the theme which is assumed to replace the whole application theme and thus affects all the controls
For points 2 and 3 there is a solution though: Grunt utility. OpenUI5 itself uses Grunt for producing builds and there is a github repository containing probably most required packages to produce minified version of the library and CSS files. However you should be aware that this is quite a low level tool based on Node.js server side JavaScript run-time.
UPDATE (22-Feb-2017): For points 2 and 3 please see Robin’s excellent blog.
Thanks again for sharing this Sergey. I’m currently looking into using Grunt to build the custom library and deploy it to HCP and or an ABAP AS. I’ll post on SCN if I find a way to do it.
Thanks Pierre, that would be helpful.
Hi Sergey,
thank you for sharing your findings. I was able to reproduce and understand part 1 of this blog series. Currently I don’t have access to an ABAP system so I cannot reproduce part 2.
But I don’t understand why NW server maps “my.custom.control.ProductRating.js” to “http://…/sap/bc/ui5_ui5/sap/zcustlib/~C2BA8580F0AD68211B5F48EF79A345B7~5/ProductRating.js“.
You didn’t define this mapping from namespace to zcustlib in a descriptor file like you similarly did on HCP in nep-app.json. Does the NW server scan the manifest.json during deployment and save the information in a table where it looks up unresolved namespaces at runtime or is there another reason for using zcustlib I don’t see?
Hi Helmut,
It appears that this is done without specific configuration in manifest. Unfortunately by this instance I have no clue of the underlying mechanism. It should be hidden inside NW logic of converting project specific url into NW specific one. As you might remember when you deploy UI5 app to SAP ABAP it requires a BSP name which is then stored in .project.json (section “deploy”) file in Web IDE, and apparently NW uses BSP name with additional GUID part to map to the root folder.
Just to add: in manifest file of the library project (section “sap.app”) you specify an id of the library, which normally is its name space (in our case “my.custom.control”), then in .project.json (which stores the Web IDE project settings) you specify the BSP name for deploying to SAP ABAP server, and most likely when deployment proceeds the SAP NW stores some map between BSP name and the project ID and that’s how it can decipher library file name and convert it to the SAP NW specific url.
Thanks Sergey for your explanation. I will test if it runs as expected probably next week.
Hello Helmut and Sergey,
What you also can do is add the control library “my.custom.control” to the dependencies (libs) of sap.ui5 in your manifest of your UI5 application. But you need to preload your control library, because it seems that the framework searches for a library-preload.json (or library-preload.js from version 1.40), without a fallback to the basic library.js file. SO the first time you would open your application you will receive the “Could not open app.”-error. But the second time you try to open your application it would open it successfully…
So that first error you can solve by using the openui5_preload grunt task for the library. Next thing that I found that can block things is that you really, really need to have a .library file in your dist folder.
Normally this way it should work, or at least is working here in my system. Maybe I need to create a blog covering these last steps?
Robin
Hi Robin,
Yes, it would be really great if you could post a blog on GRUNTing a custom library. Actually the best way would be to have some support for developing libraries in SAP Web IDE, but having a blog on GRUNT would also be a good step.
Thank you.
Hi Sergey,
You can find the steps I did on following blog .
Thank you Robin! I inserted a link to your post into this blog.
Hello, can you please tell me how to deloy a ui5 library to Fiori Launch pad | https://blogs.sap.com/2016/12/20/sapui5-custom-control-library-web-ide-development-deployment-to-hcp-and-to-on-premise-abap-repository.-part-2./ | CC-MAIN-2017-39 | en | refinedweb |
0
Hello - I'm a brand new student of programming and I'm needing help. I'm stonewalled. I've got an assignment that requires me to create a class named Dice that another program will use to roll dice.
The instructions are:
- The class must be named Dice
- A method named diceRoll() must be included
- A random number between 1 and 6 inclusive must be generated
- The random number value generated must be returned
- The face of the die must be displayed below: (description of display not included here - I've got that in my code, and it works right)
Here is my code thus far. It compiles/runs fine - I think it does what it's supposed to do, BUT... I don't know how to create or implement or include this method diceRoll(), which professor's program needs to make hers work. This part confuses me. I take this course online - the book is helpful, but I have had little or no communication or help otherwise. Thanks in advance for any ideas:
import java.util.Random; // Needed for random dice numbers. public class Dice { public static void main (String[] args) { int face; // The face of the die. // A random object for the program. Random randomNumbers = new Random(); // Die face number is set equal to a random number. face = randomNumbers.nextInt(6)+ 1; // Display results. if (face == 1) System.out.println("One:\n\n *"); else if (face == 2) System.out.println("Two:\n*\n\n *"); else if (face == 3) System.out.println("Three:\n*\n *\n *"); else if (face == 4) System.out.println("Four:\n* *\n\n* *"); else if (face == 5) System.out.println("Five:\n* *\n *\n* *"); else if (face == 6) System.out.println("Six:\n* *\n* *\n* *"); } }
Edited by mike_2000_17: Fixed formatting | https://www.daniweb.com/programming/software-development/threads/109376/beginner-help | CC-MAIN-2018-34 | en | refinedweb |
Represents a shadow constructor declaration introduced into a class by a C++11 using-declaration that names a constructor. More...
#include "clang/AST/DeclCXX.h"
Represents a shadow constructor declaration introduced into a class by a C++11 using-declaration that names a constructor.
For example:
Definition at line 3243 of file DeclCXX.h.
Definition at line 3345 of file DeclCXX.h.
References clang::Decl::getKind().
Returns
true if the constructed base class is a virtual base class subobject of this declaration's class.
Definition at line 3335 of file DeclCXX.h.
Referenced by clang::CodeGen::CodeGenTypes::inheritingCtorHasParams().
Definition at line 2656 of file DeclCXX.cpp.
Referenced by clang::Sema::BuildUsingShadowDecl().
Definition at line 2664 of file DeclCXX.cpp.
Get the constructor or constructor template in the derived class correspnding to this using shadow declaration, if it has been implicitly declared already.
Get the base class that was named in the using declaration.
This can be different for each redeclaration of this same shadow decl.
Definition at line 2668 of file DeclCXX.cpp.
References clang::NestedNameSpecifier::getAsRecordDecl(), clang::UsingDecl::getQualifier(), and clang::UsingShadowDecl::getUsingDecl().
Returns the parent of this using shadow declaration, which is the class in which this is declared.
Definition at line 3299 of file DeclCXX.h.
Referenced by clang::Sema::DefineInheritingConstructor(), and clang::Sema::findInheritingConstructor(). | http://clang.llvm.org/doxygen/classclang_1_1ConstructorUsingShadowDecl.html | CC-MAIN-2018-34 | en | refinedweb |
#include <deal.II/fe/fe_q_dg0.h>
Implementation of a scalar Lagrange finite element
Qp+DG0 that yields the finite element space of continuous, piecewise polynomials of degree
p in each coordinate direction plus the space of locally constant functions..
For more information regarding this element see: Boffi, D., et al. "Local Mass Conservation of Stokes Finite Elements." Journal of Scientific Computing (2012): 1-18.
The constructor creates a TensorProductPolynomials object that includes the tensor product of
LagrangeEquidistant polynomials of degree
p plus the locally constant function. This
TensorProductPolynomialsConst object provides all values and derivatives of the shape functions. In case a quadrature rule is given, the constructor creates a TensorProductPolynomialsConst object that includes the tensor product of
Lagrange polynomials with the support points from
points and a locally constant function.
Furthermore the constructor fills the
interface_constrains, the
prolongation (embedding) and the
restriction matrices.
The original ordering of the shape functions represented by the TensorProductPolynomialsConst is a tensor product numbering. However, the shape functions on a cell are renumbered beginning with the shape functions whose support points are at the vertices, then on the line, on the quads, and finally (for 3d) on the hexes. Finally there is a support point for the discontinuous shape function in the middle of the cell. To be explicit, these numberings are listed in the following:
1D case:
* 0---2---1 *
2D case:
* 2-------3 * | | * | 5 | * | | * 0-------1 *
3D case:
* 6-------7 6-------7 * /| | / /| * / | | / / | * / | | / / | * 4 | 8 | 4-------5 | * | 2-------3 | | 3 * | / / | | / * | / / | | / * |/ / | |/ * 0-------1 0-------1 *
The respective coordinate values of the support points of the degrees of freedom are as follows:
[ 0, 0, 0];
[ 1, 0, 0];
[ 0, 1, 0];
[ 1, 1, 0];
[ 0, 0, 1];
[ 1, 0, 1];
[ 0, 1, 1];
[ 1, 1, 1];
[1/2, 1/2, 1/2];
1D case:
* 0---2---1 *
Index 3 has the same coordinates as index 2
2D case:
* 2---7---3 * | | * 4 8 5 * | | * 0---6---1 *
Index 9 has the same coordinates as index 2
3D case:
* 6--15---7 6--15---7 * /| | / /| * 12 | 19 12 1319 * / 18 | / / | * 4 | | 4---14--5 | * | 2---11--3 | | 3 * | / / | 17 / * 16 8 9 16 | 9 * |/ / | |/ * 0---10--1 0---8---1 * * *-------* *-------* * /| | / /| * / | 23 | / 25 / | * / | | / / | * * | | *-------* | * |20 *-------* | |21 * * | / / | 22 | / * | / 24 / | | / * |/ / | |/ * *-------* *-------* *
The center vertices have number 26 and 27.
The respective coordinate values of the support points of the degrees of freedom are as follows:
[0, 0, 0];
[1, 0, 0];
[0, 1, 0];
[1, 1, 0];
[0, 0, 1];
[1, 0, 1];
[0, 1, 1];
[1, 1, 1];
[0, 1/2, 0];
[1, 1/2, 0];
[1/2, 0, 0];
[1/2, 1, 0];
[0, 1/2, 1];
[1, 1/2, 1];
[1/2, 0, 1];
[1/2, 1, 1];
[0, 0, 1/2];
[1, 0, 1/2];
[0, 1, 1/2];
[1, 1, 1/2];
[0, 1/2, 1/2];
[1, 1/2, 1/2];
[1/2, 0, 1/2];
[1/2, 1, 1/2];
[1/2, 1/2, 0];
[1/2, 1/2, 1];
[1/2, 1/2, 1/2];
[1/2, 1/2, 1/2];
1D case:
* 0--2-4-3--1 *
* 2--10-11-3 * | | * 5 14 15 7 * | 16 | * 4 12 13 6 * | | * 0--8--9--1 *
1D case:
* 0--2--3--4--1 *
Index 5 has the same coordinates as index 3
* 2--13-14-15-3 * | | * 6 22 23 24 9 * | | * 5 19 20 21 8 * | | * 4 16 17 18 7 * | | * 0--10-11-12-1 *Index 21 has the same coordinates as index 20
Definition at line 237 of file fe_q_dg0.h.
Constructor for tensor product polynomials of degree
p plus locally constant functions.
Definition at line 33 of file fe_q_dg0.cc.
Constructor for tensor product polynomials with support points
points plus locally constant functions based on a one-dimensional quadrature formula. The degree of the finite element is
points.size()-1. Note that the first point has to be 0 and the last one 1.
Definition at line 59 of file fe_q_dg0.cc.
Return a string that uniquely identifies a finite element. This class returns
FE_Q_DG0<dim>(degree), with
dim and
degree replaced by appropriate values.
Implements FiniteElement< dim, spacedim >.
Definition at line 89 of file fe_q_dg0 180 of file fe_q_dg0.cc.
Interpolate a set of scalar values, computed in the generalized support points.
Reimplemented from FiniteElement< dim, spacedim >.
Definition at line 206 of file fe_q_dg0 227 of file fe_q_dg0.cc.
Interpolate a set of vector values, computed in the generalized support points.
Reimplemented from FiniteElement< dim, spacedim >.
Definition at line 254 of file fe_q_dg0.cc.
Return the matrix interpolating from the given finite element to the present one. The size of the matrix is then
dofs_per_cell times
source.dofs_per_cell.
These matrices are only available if the source element is also a
FE_Q_DG0 element. Otherwise, an exception of type FiniteElement<dim,spacedim>::ExcInterpolationNotImplemented is thrown.
Reimplemented from FE_Q_Base< TensorProductPolynomialsConst< dim >, dim, spacedim >.
Definition at line 282 of file fe_q_dg0.cc.
This function returns
true, if the shape function
shape_index has non-zero function values somewhere on the face
face_index.
Reimplemented from FE_Q_Base< TensorProductPolynomialsConst< dim >, dim, spacedim >.
Definition at line 335 of file fe_q_dg0.cc.
Return a list of constant modes of the element. For this element, there are two constant modes despite the element is scalar: The first constant mode is all ones for the usual FE_Q basis and the second one only using the discontinuous part.
Reimplemented from FE_Q_Base< TensorProductPolynomialsConst< dim >, dim, spacedim >.
Definition at line 349 of file fe_q_dg0.cc.
clone function instead of a copy constructor.
This function is needed by the constructors of
FESystem.
Implements FiniteElement< dim, spacedim >.
Definition at line 170 of file fe_q_dg0.cc.
Return the restriction_is_additive flags. Only the last component is true.
Definition at line 310 of file fe_q_dg0.cc.
Only for internal use. Its full name is
get_dofs_per_object_vector function and it creates the
dofs_per_object vector that is needed within the constructor to be passed to the constructor of
FiniteElementData.
Definition at line 321 of file fe_q_dg0.cc. | https://dealii.org/8.5.0/doxygen/deal.II/classFE__Q__DG0.html | CC-MAIN-2018-34 | en | refinedweb |
#include "ltwrappr.h"
L_INT LBarCode::GetDuplicatedCount (nIndex)
Gets the total number of barcodes, within the class object's barcode data array, that are duplicates of the barcode at the specified index.
Index of a barcode in the class object's barcode data array. This index is zero-based. nIndex must be >= 0 and less than the total number of barcodes read using the LBarCode::Read function.
The LBarCode::Read function returns the total number of barcodes read.
Required DLLs and Libraries
For complete sample code, refer to the BarCode demo. For an example, refer to LBarCode::GetFirstDuplicated(). | https://www.leadtools.com/help/leadtools/v20/barcode/clib/lbarcode-getduplicatedcount.html | CC-MAIN-2018-34 | en | refinedweb |
J.. package alphabets;
import java.util
View Tutorial By: Nitish at 2013-01-25 04:06:38
2. Hi for Java beginners this is the right place to s
View Tutorial By: Anon at 2009-02-01 21:14:28
3. Hi Ramlak
First of all thanks for u
View Tutorial By: Vikas sharma at 2010-05-08 01:09:31
4. This tutorial is really good for beginners.It help
View Tutorial By: Prabhat Kumar at 2007-04-09 05:02:02
5. very good cod
View Tutorial By: moly at 2007-03-13 08:17:40
6. very good example and explanation. This kind of p
View Tutorial By: jagdish at 2011-07-03 15:15:02
7. How to fix this Error.....
org.apac
View Tutorial By: james prabhakaran at 2014-12-02 13:08:30
8. pls send simple to hard j2me coding
View Tutorial By: Seenu at 2008-10-29 15:19:03
9. Hii I m Ravi. This is a good example.
Here
View Tutorial By: Ravi at 2011-02-17 09:09:06
10. The best explanation....Thanq :)
View Tutorial By: Gautham at 2012-01-29 05:05:48 | http://java-samples.com/showcomment.php?commentid=34790 | CC-MAIN-2018-34 | en | refinedweb |
So you want to implement an auto-documenting API?
Project description
A Flask extension that implements Swagger support ()
What’s Swagger?
Swagger is a spec to help you document your APIs. It’s flexible and produces beautiful API documentation that can then be used to build API-explorer-type sites, much like the one at – To read more about the Swagger spec, head over to or
Git Repository and issue tracker: Documentation:
Why do I want it?
- You want your API to be easy to read.
- You want other people to be able to use your API easily.
- You’d like to build a really cool API explorer.
- It’s Friday night and your friend just ditched on milkshakes.
How do I get it?
From your favorit shell:
$ pip install flask-sillywalk
How do I use it?
I’m glad you asked. In order to use this code, you need to first instantiate a SwaggerApiRegistry, which will keep track of all your API endpoints and documentation.
Usage:
from flask import Flask from flask.ext.sillywalk import SwaggerApiRegistry, ApiParameter, ApiErrorResponse app = Flask("my_api") registry = SwaggerApiRegistry( app, baseurl="", api_version="1.0", api_descriptions={"cheese": "Operations with cheese."}) register = registry.register registerModel = registry.registerModel
Then, instead of using the “@app.route” decorator that you’re used to using with Flask, you use the “register” decorator you defined above (or “registerModel” if you’re registering a class that describes a possible API return value).
Now that we’ve got an API registry, we can register some functions. The @register decorator, when just given a path (like @app.route), will register a GET mthod with no possible parameters. In order to document a method with parameters, we can feed the @register function some parameters.
Usage:
@register("/api/v1/cheese/random") def get_random_cheese(): """Fetch a random Cheese from the database. Throws OutOfCheeseException if this is not a cheese shop.""" return htmlify(db.cheeses.random()) @register("/api/v1/cheese/<cheeseName>", parameters=[ ApiParameter( name="cheeseName", description="The name of the cheese to fetch", required=True, dataType="str", paramType="path", allowMultiple=False) ], responseMessages=[ ApiErrorResponse(400, "Sorry, we're fresh out of that cheese.") ]) def get_cheese(cheeseName): """Gets a single cheese from the database.""" return htmlify(db.cheeses.fetch(name=cheeseName))
Now, if you navigate to you should see the automatic API documentation. See documentation for all the cheese endpoints at
What’s left to do?
Well, lots, actually. This release:
- Doesn’t support XML (but do we really want to?)
- Doesn’t support the full swagger spec (e.g. “type” in data models
- Lots more. Let me know!
Project details
Release history Release notifications
Download files
Download the file for your platform. If you're not sure which to choose, learn more about installing packages. | https://pypi.org/project/Flask-Sillywalk/ | CC-MAIN-2018-34 | en | refinedweb |
Using integrated animation properties in SPAS 3.0
Article published by Pascal
July 12, 2010
One of the aims of SPAS 3.0 is to reduce the development duration of RIAs and Web sites by introducing processes in response to recurrent tasks. For this purpose, SPAS 3.0 integrates a complete windowing system since the ActionScript 1.0 version. SPAS 3.0 window classes have many helpful functionalities to improve your applications, and easily create motion effects.
In this tutorial, we will create a basic effect that can fast be adapted to all kind of applications, such as widgets or tool palettes. The effect involves an autonomous "popup" window which is able to show or hide its content when the user clicks on its title bar.
The code in this tutorial is based on SPAS 3.0 alpha 3.1 release.
We assume that you are familiar with external files programming and that you have already downloaded SPAS 3.0.
Using an Application entry point
First of all, you need to create an ActionScript class called
AnimatedPopup that extends the
Application class. Then, create a private method called "init", to be used as the application entry point. Since the alpha 2.1 version, you can use an entry point function as a parameter for the super statement of the constructor function. By doing this, you will let SPAS deal with the entry point without doing it yourself. (This hack prevents any problems that could occur when loading a child application into the main one.)
package { //--> SPAS 3.0 API import import org.flashapi.swing.*; /** * Creates a sample application that displays a popup object * which has an animated layout. */ public class AnimatedPopup extends Application { //--> Constructor public function AnimatedPopup():void { super(init); } //--> Entry point private function init():void { } } }
Designing the GUI
To create the user interface, we only need two objects: the popup and an
Image instance. The image will be used to display and manipulate an externally loaded picture. To simplify the management between both states of the popup (opened and closed), we will use a
StateManager object:
org.flashapi.swing.managers.StateManager. All these objects are declared as private variables, as shown below:
//--> GUI private var _popup:Popup; private var _manager:StateManager; private var _image:Image;
The use of the
StateManager class is probably the easiest way to manipulate different states of an object or a more complex task. But we will see how it works later; we will first focus on the GUI's design.
There are several ways to manipulate
Image instances. Here we will consider the one which is the closest to the "i
mg" tag in HTML language. This way, we use the first three parameters of the
Image constructor function, respectively the source image, its width and its height.
In our case, it is very useful, because we want to use the popup
autosize property. So, the size of the window will be adjusted according to the size of the image. Once we have created the image and the popup instances, we can add the image to the popup element list and display it.
private function init():void { _image = new Image("my_picture.jpg", 300, 326); _popup = new Popup("Double click to close"); _popup.autoSize = _popup.shadow = true; _popup.addElement(_image); _popup.display(30, 30); }
Notice that the default popup title label is "
Double click to close". It means that we invite users to double click the title bar of the popup to interact with the displayed images.
The resizing function
In the preceding part, you have discovered how it is easy to create a Graphic User Interface with SPAS 3.0! So now, let's see how to create spectacular effects with very tiny code.
At this point, what we want is to navigate between two different states. In the first one, we want to set the
height of the popup to
0 pixels, to hide the image. In the second state, we set this property to
326 pixels, to see the whole picture. Each state will see the popup titlebar's label change to indicate to the user how to proceed from the current to the next state.
This could be translated in Actionscript through the single following function:
private function setHeight(height:Number):void { _image.height = height; _popup.label = height == 0 ? "Double click to open" : "Double click to close"; }
The most OOP logical way to call this method is to define two different functions, each of which corresponds to a specific state action:
close() and
open().
private function close():void { setHeight(0); } private function open():void { setHeight(326); }
The StateManager class
If you are familiar with Actionscript coding, you have probably written a lot of similar lines of code. But what is really annoying in this process is to add the corresponding event handlers, because of the multiplication of identical actions.
With SPAS 3.0, we can use the
StateManager class, for more fun. Instead of creating several events for all states, we just add an event producer. Then we can add as many state actions as we want, by using the
addActions() method:
_manager = new StateManager(); _manager.addEventProducer(_popup, WindowEvent.TITLE_BAR_DOUBLE_CLICK); _manager.addActions(close, open);
With the
addEventProducer() method, we indicate that the popup must fire a state action each time its titlebar is double clicked by the user.
To see how the application works, write the above code at the end of the
init() function and compile your Flash movie.
Using animation properties
So now, we have got a fully functional and reactive application. But what about adding an ajax-like effect when the popup window is resized?
Let's go, and use the
autoSizeAnimated property, which is defined by the
Layout interface! Add the following code after the
StateManager statement, and compile the animation again to see what happens:
_popup.layout.autoSizeAnimated = true;
Actually, the
Layout
autoSizeAnimated property is an automatic process implemented by some layouts to make spectacular effects with a minimum of work.
(Notice that this property does not work at the moment if it is combined with the
Layout
animated property.)
And here is the complete class we wrote all along this tutorial:
package { import org.flashapi.swing.*; import org.flashapi.swing.event.*; import org.flashapi.swing.managers.*; public class AnimatedPopup extends Application { public function AnimatedPopup():void { super(init); } private var _popup:Popup; private var _manager:StateManager; private var _image:Image; private function init():void { _image = new Image("my_picture.jpg", 300, 326); _popup = new Popup("Double click to close"); _popup.autoSize = _popup.shadow = true; _popup.addElement(_image); _popup.display(30, 30); _manager = new StateManager(); _manager.addEventProducer(_popup, WindowEvent.TITLE_BAR_DOUBLE_CLICK); _manager.addActions(close, open); _popup.layout.autoSizeAnimated = true; } private function close():void { setHeight(0); } private function open():void { setHeight(326); } private function setHeight(height:Number):void { _image.height = height; _popup.label = height == 0 ? "Double click to open" : "Double click to close"; } } }
Result
Share this article:
There are no comments yet for this article. | http://www.flashapi.org/tutorials/?url=using-integrated-animation-properties&PHPSESSID=fb9f7ec637cdee341fe775aa5471c7fb | CC-MAIN-2018-34 | en | refinedweb |
Tokens from Complex Objects¶
A very common setup is to have your users information (usernames,
passwords, roles, etc) stored in a database. Now, lets pretend that we want to
create an access tokens where the tokens identity is a username, and we also want
to store a users roles as an additional claim in the token. We can do
this with the
user_claims_loader()
decorator, discussed in the previous section. However, if we pass the username
to the
user_claims_loader(), we would end
up needing to query this user from the database two times. The first time would
be when login endpoint is hit and we need to verify a username and password.
The second time would be in the
user_claims_loader()
function, because we need to query the roles for this user. This isn’t a huge
deal, but obviously it could be more efficient.
This extension provides the ability to pass any object to the
create_access_token() function, which will then be
passed as is to the
user_claims_loader().
This allows us access the database only once, but introduces a new
issue that needs to be addressed. We still need to pull the username
out of the object, so that we can have the username be the identity for the
new token. We have a second decorator we can use for this,
user_identity_loader(), which lets you
take any object passed in to
create_access_token()
and return a json serializable identity from that object.
Here is an example of this in action:
from flask import Flask, jsonify, request from flask_jwt_extended import ( JWTManager, jwt_required, create_access_token, get_jwt_identity, get_jwt_claims ) app = Flask(__name__) app.config['JWT_SECRET_KEY'] = 'super-secret' # Change this! jwt = JWTManager(app) # This is an example of a complex object that we could build # a JWT from. In practice, this will likely be something # like a SQLAlchemy instance. class UserObject: def __init__(self, username, roles): self.username = username self.roles = roles # Create a function that will be called whenever create_access_token # is used. It will take whatever object is passed into the # create_access_token method, and lets us define what custom claims # should be added to the access token. @jwt.user_claims_loader def add_claims_to_access_token(user): return {'roles': user.roles} # Create a function that will be called whenever create_access_token # is used. It will take whatever object is passed into the # create_access_token method, and lets us define what the identity # of the access token should be. @jwt.user_identity_loader def user_identity_lookup(user): return user.username @app.route('/login', methods=['POST']) def login(): username = request.json.get('username', None) password = request.json.get('password', None) if username != 'test' or password != 'test': return jsonify({"msg": "Bad username or password"}), 401 # Create an example UserObject user = UserObject(username='test', roles=['foo', 'bar']) # We can now pass this complex object directly to the # create_access_token method. This will allow us to access # the properties of this object in the user_claims_loader # function, and get the identity of this object from the # user_identity_loader function. access_token = create_access_token(identity=user) ret = {'access_token': access_token} return jsonify(ret), 200 @app.route('/protected', methods=['GET']) @jwt_required def protected(): ret = { 'current_identity': get_jwt_identity(), # test 'current_roles': get_jwt_claims() # ['foo', 'bar'] } return jsonify(ret), 200 if __name__ == '__main__': app.run() | http://flask-jwt-extended.readthedocs.io/en/latest/tokens_from_complex_object.html | CC-MAIN-2018-34 | en | refinedweb |
Python vs Java – Who Will Conquer 2019?
Python Vs Java – The hottest battle of the era. Every beginner want to know which programming language will have a bright future? According to statistics, Java is losing its charm and Python is rising. But, no one will tell you which one is beneficial. In this blog, we will discuss the differences between Java and Python and let you decide which one is more useful.
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Python Vs Java – A Battle for the Best
Let’s deep dive into the differences.
1. Hello World Example
To rigorously compare Python and Java, we first compare the first program in any programming language- to print “Hello World”.
- Java
public class HelloWorld { public static void main(String[] args) { System.out.println("Hello World"); } }
- Python
Now, let’s try printing the same thing in Python.
print(“Hello World”)
As you can see, what we could do with 7 lines of code in Java, we can do with 1 line in Python.
Let’s further discuss parts of this program and other things.
2. Syntax
A striking characteristic of Python is simple python syntax. Let’s see the syntax difference between Python and Java.
2.1 Semicolon
Python statements do not need a semicolon to end, thanks to its syntax.
>>> x=7 >>> x=7;
But it is possible to append it. However, if you miss a semicolon in Java, it throws an error.
class one { public static void main (String[] args) { int x=7; System.out.println(x) } }
Compilation error #stdin compilation error #stdout 0.09s 27828KB
Main.java:10: error: ‘;’ expected
System.out.println(x) ^
1 error
2.2 Curly Braces and Indentation
The major factor of Python Vs Java.
>>> if 2>1: print("Greater")
Greater
This code would break if we added curly braces.
>>> if 2>1: {
SyntaxError: expected an indented block
Now, let’s see if we can skip the curly braces and indentation in Java.
class one { public static void main (String[] args) { if(2>1) System.out.println("2"); } }
Success #stdin #stdout 0.07s 27792KB
2
Here, we could skip the braces because it’s a single-line if-statement. Indentation isn’t an issue here. This is because when we have only one statement, we don’t need to define a block. And if we have a block, we define it using curly braces. Hence, whether we indent the code or not, it makes no difference. Let’s try that with a block of code for the if-statement.
class one { public static void main (String[] args) { if(2<1) System.out.println("2"); System.out.println("Lesser"); } }.
2.3 Parentheses
Starting Python 3.x, a set of parentheses is a must only for the print statement. All other statements will run with or without it.
>>> print("Hello")
Hello
>>> print "Hello"
SyntaxError: Missing parentheses in call to ‘print’
This isn’t the same as Java, where you must use parentheses.
2.4 Comments
Comments are lines that are ignored by the interpreter. Java supports multiline comments, but Python does not. The following are comments in Java.
//This is a single-line comment /*This is a multiline comment Yes it is*/
Now, let’s see what a comment looks like in Python.
>>> #This is a comment
Here, documentation comments can be used in the beginning of a function’s body to explain what
it does. These are declared using triple quotes (“””).
>>> """ This is a docstring """ '\n\tThis is a docstring\n'
These were the syntax comparison in Python Vs Java, let’s discuss more.
3. Dynamically Typed
One of the major differences is that Python is dynamically-typed. This means that we don’t need to declare the type of the variable, it is assumed at run-time. This is called Duck Typing. If it looks like a duck, it must be a duck, mustn’t it?
>>> age=22
You could reassign it to hold a string, and it wouldn’t bother.
>>> age='testing'
In Java, however, you must declare the type of data, and you need to explicitly cast it to a different type when needed. A type like int can be casted into a float, though, because int has a shallower range.
class one { public static void main (String[] args) { int x=10; float z; z=(float)x; System.out.println(z); } }
Success #stdin #stdout 0.09s 27788KB
10.0
However then, at runtime, the Python interpreter must find out the types of variables used. Thus, it must work harder at runtime.
Java, as we see it, is statically-typed. Now if you declare an int an assign a string to it, it throws a type exception.
class one { public static void main (String[] args) { int x=10; x="Hello"; } }
Compilation error #stdin compilation error #stdout 0.09s 27920KB
Main.java:12: error: incompatible types: String cannot be converted to int
x="Hello"; ^
1 error
4. Verbosity/ Simplicity
Attributed to its simple syntax, a Python program is typically 3-5 times shorter than its counterpart in Java. As we have seen earlier, to print “Hello World” to the screen, you need to write a lot of code in Java. We do the same thing in Python in just one statement. Hence, coding in Python raises programmers’ productivity because they need to write only so much code needed. It is concise.
To prove this, we’ll try to swap two variables, without using a third, in these two languages. Let’s begin with Java.
class one { public static void main (String[] args) { int x=10,y=20; x=x+y; y=x-y; x=x-y; System.out.println(x+" "+y); } }
Success #stdin #stdout 0.1s 27660KB
20 10
Now, let’s do the same in Python.
>>> a,b=2,3 >>> a,b=b,a >>> a,b
(3, 2)
As you can see here, we only needed one statement for swapping variables a and b. The statement before it is for assigning their values and the one after is for printing them out to verify that swapping has been performed. This is a major factor of Python vs Java.
5. Speed
When it comes to speed, Java is the winner. Since Python is interpreted, we expect them to run slower than their counterparts in Java. They are also slower because the types are assumed at run time. This is extra work for the interpreter at runtime. The interpreter follows REPL (Read Evaluate Print Loop). Also, the IDLE has built-in syntax highlighting, and to get the previous and next commands, we press Alt+p and Alt+n respectively.
However, they also are quicker to develop, thanks to Python’s brevity. Therefore, in situations where speed is not an issue, you may go with Python, for the benefits it offers more than nullify its speed limitations. However, in projects where speed is the main component, you should go for Java. An example of such a project is where you may need to retrieve data from a database. So if you ask Python Vs Java as far as speed is concerned, Java wins.
6. Portability
Both Python and Java are highly portable languages. But due to the extreme popularity of Java, it wins this battle. The JVM (Java Virtual Machine) can be found almost everywhere. In the Python Vs Java war of Portability, Java wins.
7. Database Access
Like we’ve always said, Python’s database access layers are weaker than Java’s JDBC (Java DataBase Connectivity). This is why it isn’t used in enterprises rarely use it in critical database applications.
8. Interpreted
With tools like IDLE, you can also interpret Python instead of compiling it. While this reduces the program length and boosts productivity, it also results in slower overall execution.
9. Easy to Use
Now because of its simplicity and shorter code, and because it is dynamically-typed, Python is easy to pick up. If you’re just stepping into the world of programming, beginning with Python is a good choice. Not only is it easy to code, but it is also easy to understand. Readability is another advantage. However, this isn’t the same as Java. Because it is so verbose, it takes some time to really get used to it.
10. Popularity and Community
If we consider the popularity and community factor for Python vs Java, we see that for the past few decades, Java has been the 2nd most popular language (TIOBE Index). It has been here since 1995 and has been the ‘Language of the Year’ in the years 2005 and 2015. It works on a multitude of devices- even refrigerators and toasters.
Python, in the last few years, has been in the top 3, and was titled ‘Language of the Year’ in years 2007, 2010, and 2018. Python has been here since 1991. Can we just say it is the easiest to learn? It is a great fit as an introductory programming language in schools. Python is equally versatile with applications ranging from data science and machine learning to web development and developing for Raspberry Pi.
While Java has one large corporate sponsor- Oracle, Python is open-source (CPython) and observes distributed support.
11. Use Cases
Python: Data Science, Machine Learning, Artificial Intelligence, and Robotics, Websites, Games, Computer Vision (Facilities like face-detection and color-detection), Web Scraping (Harvesting data from websites), Data Analysis, Automating web browsers, Scripting, Scientific Computing
Java: Application servers, Web applications, Unit tests, Mobile applications, Desktop applications, Enterprise applications, Scientific applications, Web and Application Servers, Web Services, Cloud-based applications, IoT, Big Data Analysis, Games
12. Best for
While Python is best for Data Science, AI, and Machine Learning, Java does best with embedded and cross-platform applications.
13. Frameworks
Python: Django, web2py, Flask, Bottle, Pyramid, Pylons, Tornado, TurboGears, CherryPy, Twisted
Java: Spring, Hibernate, Struts, JSF (Java Server Faces), GWT (Google Web Toolkit), Play!, Vaadin, Grails, Wicket, Vert.x
14. Preferability for Machine Learning and Data Science
Python is easier to learn and has simpler syntax than Java. It is better for number crunching, whereas Java is better for general programming. Both have powerful ML libraries- Python has PyTorch, TensorFlow, scikit-learn, matplotlib, and Seaborn, and Java has Weka, JavaML, MLlib, and Deeplearning4j
Similarities of Python and Java
Besides differences, there are some similarities between Python and Java:
- In both languages, almost everything is an object
- Both offer cross-platform support
- Strings are immutable in both- Python and Java
- Both ship with large, powerful standard libraries
- Both are compiled to bytecode that runs on virtual machines
This was all about the difference between Python vs Java Tutorial.
Summary
So, after all, that we’ve discussed here in Python vs Java Tutorial, we come to conclude that both languages have their own benefits. It really is up to you to choose one for your project. While Python is simple and concise, Java is fast and more portable. While Python is dynamically-typed, Java is statically-typed. Both are powerful in their own realms, but we want to know which one you prefer. Furthermore, if you have any query/question, feel free to share with us!
So now you know which out of java and python is best for your project, install python on Windows if you are willing to go ahead with Python.
Java has greatly changed since version 1.8. With lambdas, streams, default methods, code has become less verbose, concise and started favoring parallel computation. With no second thought, Java / Scala would be my choice.
Hi G.Sridhar,
Thank you for sharing such a nice piece of information on our Python vs Java Tutorial. We have a series of Scala tutorial as well refer them too.
Keep learning and keep visiting Data Flair
You have tones of false claims on comparison. How do you compare speed of two languages? Under what circumstances? Just another blowing click bait
Hey Osman,
Python, as mentioned above, is an interpreter based language. An interpreter based language being slower than a compiler based one is the first conclusion that anyone would deduce. Java excels at speed when it is compared with Python. This is the main reason as to why it is preferred as a server-side computing language. Java is also statically typed as opposed to Python which is dynamically typed. There are various evidences and benchmarks that bolster the claim that Java executes programs faster than Python.
Hope, it helps! | https://data-flair.training/blogs/python-vs-java/ | CC-MAIN-2020-16 | en | refinedweb |
Get objects highlighted property, is it possible?
On 23/06/2013 at 08:30, xxxxxxxx wrote:
Hi,
I hope someone can help me with my tiny problem if its not to much bother?
I've been trying to find a way to perform a similar function to the command line where by you can drag an objects property to the command line and it will give the details of that property. Is there any way to replicate this so that a user could highlight an objects property and a script act upon that highlighted property of that selected object?
I wasn't sure if it was possible so tried to go down a different route and it looks like I may be going down the wrong route. I've tried to make a text field input that a user can copy and paste from the command line into the text field and then get a script to work on that, but with no luck.
Has any one had any luck with this or a similar approach, I'd appreciate any ideas?
import c4d import re def main() : obj = op.GetObject() ud = op[c4d.ID_USERDATA,2] pat = re.compile(r'[.*]') strip_ud = pat.search(ud).group(0) udata = "obj"+strip_ud # This wont work as its a string not the actual objects property udata = 500 # Below is what I'm trying to replicate. #obj[c4d.PRIM_TUBE_HEIGHT] = 500
Here is a link to the scene file to:
Thanks guys
On 23/06/2013 at 09:27, xxxxxxxx wrote:
i am not sure what you are trying to do. generally said it is impossible to access the
attribute manger in python. i think it is even not possible in cpp to get the currently
selected description element in the attribute manager. however:
1. the c4d module contains a dict with almost all ids. in some rare places there are
black spots, do not ask me why.
2. the pythonic way to do what you want are the getattr/setattr/getitem/setitem
methods. The bracket syntax is just a wrapper for those important methods. it depends
on the class you invoke __get/setitem__on what it does expect as the key parameter,
for a GeListNode it expects a DescID, so we have to get that id from the c4d dict first.
You can also invoke the __get/setitem__ methods directly on an objects BaseContainer,
~~where it does accept a key string as the key value. ~~Hm, just looked it up to be sure and
its turns out that a BaseContainer does not not accept strings either, i must have mixed
something up there, not sure why i thought so. So you have to stick with the DescIDs for
all cases.
import c4d ID_NAME = getattr(c4d, 'ID_BASELIST_NAME') ID_RELPOS = getattr(c4d, 'ID_BASEOBJECT_REL_POSITION') def main() : if isinstance(op, c4d.BaseObject) : oldname = op.__getitem__(ID_NAME) print oldname op.__setitem__(ID_NAME, oldname + '_x') oldpos = op.__getitem__(ID_RELPOS) print oldpos if isinstance(oldpos, c4d.Vector) : oldpos += c4d.Vector(0,100,0) op.__setitem__(ID_RELPOS, oldpos) c4d.EventAdd() if __name__=='__main__': main()
On 26/06/2013 at 07:51, xxxxxxxx wrote:
Thanks littledevil. I suspected it wasn't possible. I'll have a look at the method you suggested and see how I get on, so thanks for that.
On 01/07/2013 at 15:08, xxxxxxxx wrote:
I recently saw Nitro4D's video for his MagicLoadTexture plugin and saw something very interesting at around the 6 minuet mark where by he adds some input boxes to his lay out and then is able to drag material properties into these input boxes. Does any one have any idea how he may have done this? This is the sort of thing I was looking for I think. It would be great to drag in what ever attribute you would like (with in reason) and then act upon it, an objects position for example. Any ideas?
Here is a link to the video:
On 01/07/2013 at 15:18, xxxxxxxx wrote:
you mean the textbox holding c4d.material.someid ? looks like a multilineedit with the python flag.
On 03/07/2013 at 01:53, xxxxxxxx wrote:
Ah, thanks I'll I have a look into that and hopefully getting it working. Thanks littledevil | https://plugincafe.maxon.net/topic/7283/8443_get-objects-highlighted-property-is-it-possible | CC-MAIN-2020-16 | en | refinedweb |
Request For Commits – Episode #1
Open Source, Then and Now (Part 1)
with Karl Fogel, author of Producing Open Source Software
Nadia Eghbal and Mikeal Rogers kick off Season 1 of Request For Commits with a two part conversation with Karl Fogel — a software developer who has been active in open source since its inception.
Karl served on the board of the Open Source Initiative, which coined the term “open source”, and helped write Subversion, a popular version control system that predates Git. Karl also wrote a popular book on managing open source projects called Producing Open Source Software. He’s currently a partner at Open Tech Strategies, a firm that helps major organizations use open source to achieve their goals.
- Read Karl Fogel’s book — Producing Open Source Software
- Listen to part 2 with Karl Fogel as we continue this conversation.
Transcript
I’m Nadia Eghbal.
And I’m Mikeal Rogers.
On today’s show, Mikeal and I talked with Karl Fogel, author of Producing Open Source Software. Karl served on the board of the Open Source Initiative, which coined the term ‘open source’ and helped write Subversion. He’s currently a partner at Open Tech Strategies, helping major organizations use open source to achieve their goals.
Our focus on today’s episode with Karl was about what has changed in open source since he first published his book ten years ago. We talked about the influence of Git and GitHub, and how they’ve changed both development workflows and our culture.
We also talked about changes in the wider perception of open source, whether open source has truly won, and the challenges that still remain.
So back in 2006 I started working at the Open Source Applications foundation on the Chandler Project, and I remember we had to kind of put together a governance policy and how do we manage an open source project, how do we do it openly, and basically your book kind of got slapped on everybody’s desk. [laughter] The Producing Open Source Software first edition, and it was like “This is how you run open source projects.”
Wow, that’s really nice to hear, thank you.
And it was… Especially at that time it was an amazing guide, and I know from talking with Jacob Kaplan-Moss that the Django project did something similar, as well. I’m very curious how you got to write that book and what preceded it. It’s produced by O’Reilly, right?
Yes.
I’m curious why O’Reilly wanted to do something… It’s very deep and very nerdy, so…
Yeah, actually I wanna take a quick second to give a shout out to O’Reilly because… I mean, that was never a book that was gonna be a bestseller, and they sort of knew that from the beginning, and they not only decided to produce it anyway, they gave me a very good editor, Andy Oram, who made a lot of contributions to the book in terms of shaping it and giving good feedback. And they let me publish it under a free license, which to a publisher, that’s a pretty big move, and it’s not something that they do with all their books. So I really appreciated the support I got from them.
So the answer to your main question there I’m afraid is pure luck. I really think that in the early 2000s, 2005-2006 the time was ripe for some kind of long-form guide to the social and community management aspects of open source to come out, and my book just happened to come out. If someone else had written a long-form guide, then… You know, it’s like in the early days of physics - if you just happen to be the first person to think of calculus, you’ll get all this credit; but there were probably ten people who thought of it, it’s just that someone published this first.
So yeah, I just got really lucky with the timing. And the way that I was motivated to write it, that O’Reilly had contacted me about doing a Subversion book… I was coming off five or six years as a founding developer in the Subversion project and it had been my full-time job, and I’d gone from being mostly a programmer and sort of project technical - not necessarily technical lead, but technical arbiter or technical cheerleader in some sense, to more and more community manager. I mean, I was still doing coding, but a lot of my time was spent on just organizing and coordinating the work of others and interjecting what I felt were the appropriate noises in certain contentious discussion threads and things like that.
[00:04:09.07] So when it came time to write a Subversion book, I had already written a book, I knew folks at O’Reilly, and they said “Would you like to be one of the authors?” There were a couple other Subversion developers that I worked with who were also interested in writing, and we had all agreed that we would co-author it.
Then as I started to write, I really let down my co-authors. I said, “Hey, folks, I’m really sorry. I don’t wanna write another technical manual. I’ve already done that once. You folks go do it, it’s gonna be great.” And I wrote the introduction and they wrote a wonderful book that became one of O’Reilly’s better sellers and is still quite popular.
So I thought, “Well, what was it that I wanted to write if that wasn’t the book?” and I realized the book I wanted to write was not about Subversion the software, it was about the running of a Subversion project, and about open source projects in general - Subversion wasn’t the only one that I was involved in. So I went back to O’Reilly and I said very meekly, “Could I write this other book instead? What do you think of that?” and they said yes. So I sort of backed into it… I was forced into the realization that I wanted to write this book through trying to write another book and failing.
Was that a popular view back then? Like, when you said that you wanted to write this non-technical, more management-focused book around open source, were people like “Why?”
Let me cast back my memory… No, but then again, the people that I talked with - that’s a very biased sample, right? Most people were encouraging, and if they were mystified as to why I wanted to write this, they hid it very well and were nothing but encouraging. Then it took a little bit longer to write than I thought, and people were asking “How’s it going?” and I’d always give the same answer, like “How’s your book going?” “Never ask. Thank you.” [laughter] No one ever listened to that, they would just ask the next time. But eventually it got done.
I think there was, among people involved in open source. For example, the role of community manager was already a title you started to see people having. You started to see a phenomenon where the coordinating people, the people doing that community management and projects were no longer also the most technically sharp people. I was definitely not the best programmer in the Subversion project; I could think of a lot of names - I’ve probably even forgotten some names of people who I just think are better coders than I am, who were working on that project.
And that was true across a lot of open source projects. I could see that the people who were doing technical and community work together were not the Linus Torvalds model - and Linus Torvalds isn’t by any means a typical example… The Linux kernel in general is not a typical example of how open source projects have ever operated. It’s been its kind of own weird, unique thing for a long time. But one thing you can say about it is that the leader of the project is also one of the best programmers in the project. Linus is a very technically sharp person. But that was not the case in a lot of open source projects, and that to me seemed like a clue that, “Okay, something’s happening here where open source is maturing to the point where different sets of skills are needed”, and you’ve got these crossover people who are often founding members of a project and active in coding and other technical tasks, but their main focus, their main energy is going to the social health and the overall community health of the project.
[00:07:45.19] I wasn’t the only person sensing that. A lot of people seemed to already understand the topic of the book before I explained it to them.
For that first book, I mean, you came up through the ’90s open source scene and were clearly doing a lot of community work on the Subversion project - did you write it mostly just from your own experiences and memory, or did you go through a phase of research and reaching out to other projects?
That’s a really good question. Yeah, I researched other projects. I did rely a lot of my own experiences, which were somewhat broad; I had worked on a lot of projects by that point. But I was worried that I would be biased, and particularly towards backend systems projects, because I was a C programmer, I didn’t do a huge amount of graphical user interface programming or stuff like that. Web programming was kind of new then, but I still hadn’t done a lot of it. So I deliberately sought out some other projects to talk to and people were very generous with their time. I think I listed them all, either in the acknowledgments or in the running text or the footnote. So not all those were projects that I worked in, they were just places where people were willing to be informants.
Interesting. You mentioned that people were starting to come around and you were starting to see community manager as a title, but I do feel like the book addressed something and reset people’s expectations about how open source projects run. It did bring a lot of this community stuff and not everything being purely technical to the forefront. If there was one presumption that projects had at the time, that the book was meant to address - is there one that you can point at? Or any kind of general stories that you might have heard about shifts in people’s… What I really wanna get at is, people’s conception of open source had been this pure meritocracy, pure technical side of things, right? Not a lot had been done in a formal way to address the role of people and people management, and processes and barriers to entry until your book, as far as I know.
I think I get the question you’re asking, and it’s a good one. I’ve never really thought of the book as addressing a sort of as yet unacknowledged need, but I guess in a way it was. The observation I had at the time in Subversion, and then as I started to talk to people in other projects I realized it was just as true for them as it was for subversion, was that there’s no such thing as a structureless meritocracy, and there’s not such thing as a structureless community. We’re all heard of the famous essay The Tyranny Of Structurelessness, in which the author points out that if you think you have a structureless organization, what you really have is an organization where the rules are not clear and people with certain kind of personalities end up dominating by sometimes vicious or deceptive means. And that has certainly been the case in some open source projects. I don’t wanna name names, but we could probably all think of some.
What I saw on Subversion was that managing a bunch of people who were not all under one management hierarchy, like they were coming from different companies, and some of them were true volunteers in the sense that there was no way in which they were being paid for their time, or only very indirectly, but a lot of them were being paid for it and they had their own priorities, and to make that scene work and to have the project make identifiable progress, you had to broker compromise; you had to convince people like, “Okay, this feature that you want needs some more design work, and the community has to accept it. That means it’s not gonna be done in time for this upcoming release, but we don’t wanna delay the release because there’s another thing that this programmer or this company wants needs to be in that release and they’re depending on it. And by the way, if you get them on your side by cooperating now, he’ll be much more likely to review your changes and design when your stuff is ready.” Things like that. Making sure that the right people meet or talk at in-person events. Occasional backchannel communications to ask someone to be a little bit less harsh toward a new developer who is showing promise or is perhaps representing an important organization that is going to bring more development energy to the project, but we need to not offend their first person who comes in, who is maybe not leading with his best code; it sometimes happens.
[00:12:17.09] There were all sorts of stuff that had to be done that was not necessarily visible from just watching the public mailing list. So the book was basically - I realize I’m giving a long answer, you should feel free to edit this down, by the way… Now I’m trying to be a little less verbose…
No, this is perfect.
Okay, I’m glad. [laughs] I guess the thing the book was meant to address was you get a lot of programmers who land in open source somehow, they find themselves running projects or occupying positions of influence, and both because no one has ever said it, and because it’s not visible from the public activity on the project - or not entirely visible - and because there is a predisposition among programmers to be less aware of social things; statistically speaking, programmers I think are somewhat less socially adept people than most people. Obviously, there are exceptions to that, but I think it’s a broad categorization that is statically true. So for all of those reasons, I wanted there to be a document that said, “Hey, you need to start thinking about this as a system. You need to start thinking about the project in the same way you think about the codebase. Parts need to work together, and you need to pay attention to how people feel and to their long-term interest, and you’ve gotta put yourself in their shoes. Here’s a rough guide to doing that.”
That’s what I was thinking when I was writing the book, and I never really articulated that until you asked the question, but I’m pretty sure that’s more or less what I was thinking.
Yeah, I mean, we’re still struggling with that today. [laughs] We’re talking in the past tense because the book came out ten years ago, but I’m still struggling to get people to recognize that today…
Well, let’s go right to the controversial stuff. The Linux kernel project is famous for kind of having a toxic atmosphere, right? And Linus has basically said that he equates the thing that most of us call toxicity with meritocracy. In other words, the kinds of people who write the kinds of code that he wants to end up in a Linux kernel are the kinds of people who flourish in the atmosphere he has set up.
Maybe that’s actually true, but I just don’t think the Linux kernel project has run the experiment of trying to… Forking the project and running a nice version, where everyone is welcomed warmly and not insulted personally by a charismatic leader, in which they can see whether that theory is actually true.
Right. I was actually not even thinking about projects that are more than ten years old, but even projects that start today struggle with this. Just acknowledging that soft skills matter and that somebody needs to pick up this community work.
I think it’s interesting that you said that you wrote the book in 2005, around this time when you felt like people were starting to notice and care about the need for skills beyond coding, but I feel like that’s almost what people would say about right now too, so I wonder if anything’s even changed in ten years or not.
Well, just imagine how much worse things would be if we hadn’t all been through that. [laughter]
Yeah.
You never have an alternate universe in which to run experiment, unfortunately. But I think it will always be true, because the startup costs in open source are so low - although that’s changing a little bit, and we can talk about that later - so that the people who start projects, they’ll just land in it coming from a technical interest. They’re not starting out by thinking of soft skills, so the projects are always launched in a way that’s sort of biased towards a certain kind of culture, and then they end up having to correct toward a more socially functioning culture, even though that imposes a small amount of overhead on the technical running of the project.
[00:16:17.13] And if it’s a useful project and people are like “Well, I’m gonna use it”… Or even if it’s not useful, but it’s just kind of a legacy being used, it’s like, what incentive is there really? I think it’s still very hard to tie together, and in some cases you can tie together the health of a project with its popularity, but sometimes it’s a popular project and it’s just not that kind of place.
Yeah… I can only make anecdotal studies there. One example is the LibreOffice project - it has really gone through a great deal of trouble to be welcoming to developers and to make their initial development process easier. Building a project is now way, way easier than it used to be; they’ve just really sunk a lot of time into making it easy to compile it from source, to welcome new developers. I think that’s having a good effect, but how do you know how popular or how successful the project would be without that? You just don’t.
You mentioned that you’ve released it under Creative Commons license, and I saw that you’ve actually kind of kept it a little bit up to date and you’ve kind of pushed small changes to it over time, but in 2013 you decided to actually do a full new edition of the book. What precipitated the need for an entire new edition, rather than just adjustments?
A few things. One, the adjustments that I had been doing in the years from 2006 roughly to 2013, they weren’t that trivial. I mean, there were a lot of small scale changes that went in. I think most sections of the book got touched, some of them pretty heavily, but I was never thinking of it as a full rewrite. And then it was really partly my own feeling about certain things that were out of that and partly feedback I was getting from other people. One thing everyone noticed - and I noticed too, because I also use Git for my coding work, although I use Subversion for non-coding version control - was that all the examples used Subversion, which was totally the right thing to do in 2005, because that was the thing that you stored your open source code in, but it just wasn’t by 2013; Git was the obvious answer, and frankly even though the site itself is not open source, GitHub was clearly the thing to use. For example, most active open source code is just on GitHub, and if the book doesn’t acknowledge that fact, then it’s just not reflecting reality and it’s not giving people the smoothest entry into open source that they can have.
So one obvious thing was the revamping of all the examples to use Git instead of Subversion and to talk about GitHub. And also in general, the project hosting situation had changed. I’m sorry, I just don’t consider SourceForge a thing anymore. [laughter] So many ads, too much visual noise, not compelling enough functionality, and that’s despite the fact that the SourceForge platform itself finally went open source, as the Allure project - which is great, I’d love to be using it, but I’m afraid I just have a much better experience with GitHub and Git, so that’s what I use.
So the recommendations about how to host projects really needed to change to be oriented more around the Git-based universe and to at least acknowledge and recommend GitHub, while acknowledging that it itself is not open source… Although I hope that they see a grand strategic vision whereby opening up their actual platform makes sense some day; I think that the real secret sauce there is the dev ops, it’s not the code, so I hope they do that someday.
[00:19:53.06] The other thing that changed kind of in a big way was what I think of as the slow rise of business-to-business open source, which is… The old cliché was “Open source always starts when some individual programmer needs to scratch an itch”; she needs to analyze her log files better, so she writes a log analyzer and then she realizes that there are other sysadmins who need to do the same, so they start collaborating, and now you’ve got an open source log analyzer, and she’s the de facto leader of this new open source project. Well, that did happen a lot, but now you have things like Android. You have businesses putting out open source codebases like TensorFlow. I don’t mean to pick on Google examples only, it’s just that those are the first things that come to mind, but Facebook also does this, Hewlett Packard does it… Lots of companies are releasing open source projects which are - I guess you could call them it’s a corporation scratching a corporation’s itch, but it is not a case of an individual developer; it’s a management move, it is done for strategic reasons which they can articulate to themselves and sometimes they also articulate to the world.
And I thought that the rise of that kind kind of project needed to be covered better, and that that was a trend that if the book could explain it better to other managers in tech or tech-related companies, that perhaps it would encourage some of them to join that trend.
And sorry, I’m realizing that there’s one more component to the answer - the other thing that changed was that I expected governments to be doing more open source by 2013 than they were, and I had at that point been very active in trying to help some government agencies launch technical products as open source, because they were gonna need that technology anyway. It’s taxpayer-funded, why not make it open source? And they were just really culturally not suited to it. There were just many, many things about the way governments do technology development, in the way they do procurement, in the way they tend to be risk-averse to the exclusion of many other considerations, really made open source an uphill struggle for them, and I wanted the book to talk a lot more about that, because I wanted it to be something that government decision-makers could circulate and use as a way to reassure themselves that they could do open source and that it could be successful, and that they didn’t have to look at it as a risky move.
So there were some new trends that I wanted to cover and there were some new goals that I had for the book, and they just required ground-up reorganization and revamp.
Wow, that’s great. We’re gonna take a short break and when we come back Karl’s gonna get into how GitHub has changed the open source landscape.
We’re back with Karl Fogel. Karl, in your mind, what have Git and GitHub changed about open source today? What are the biggest shifts that happened from the Subversion Apache days to now?
[00:23:44.29] Well, so I might have to ask you for help answering this, because I wonder if I was so comfortable with old tools that maybe I was blind to something that was difficult about them. I didn’t feel like GitHub changed the culture tremendously except in the sense that Twitter changed the culture of the internet, which is to say it gave everyone an agreed-on namespace. Right now Twitter is essentially the “Hey, if you have an internet username, it’s whatever your username is on Twitter. That’s your handle now.” And in open source your GitHub handle, which for many people is the same as their Twitter handle, that’s like your chief identifier. And it’s not a completely unified namespace and there are plenty of projects that host in other places and many developers contribute to projects that are hosted in places other than GitHub… But it is sort of a unified namespace.
If you have an open source project and you don’t have the project name somewhere on GitHub, someone else is sure to take it for their fork, right? So you’ve gotta get that real estate even if you’re not hosting there.
But I think the way GitHub wants to think about it is that they made it a lot easier for people to track sources, to make lightweight quick, so-called ‘drive-by contributions’ and to maintain what used to be called vendor branches, that is to say permanent non-hostile forks; internal, or sort of feature forks that are maintained in parallel with the project, where the upstream isn’t going to ever take the patches, but they’d have otherwise no particular animosity toward the changes, or are even willing to make some adjustments so that the people who need to maintain that separate branch can do so.
So I think their goal was to make all that stuff easier, and also to make gazillions of dollars, which I’m happy to see they’re doing. And I think that it is part of GitHub’s self-identity - for the executive and upper management team, it’s part of their self-identity to think of themselves as supporting open source, that they are doing good for open source. And as I said, I always remember that the platform itself is not open source, but that aside, I think in many ways it’s true, they do a lot of things to support open source.
The moves that they made to give technical support and kind of a little nudge to projects to get real open source licenses in their repositories was a really helpful thing. Nowadays most active open source projects on GitHub do have a license file, and that’s partly because GitHub made a push to help that happen, and they’ve done a lot to support getting open source into government agencies and things like that. So I think they had sort of cultural motivations, as well as technical and financial motivations.
So has it changed the culture of open source? That’s the thing, I’m not really sure it was all that hard to contribute to an open source project before GitHub. Maybe that’s because my specialty was working on one of the tools that is the main part of the contribution workflow, with the version control tools; I worked on CVS, which was the main open source version control system in the network on Subversion, which was for a while the main open source version control system. So if I wanted to make a drive-by contribution to some other project, of course I never had any problem doing it, because the version control tool is probably something I hacked on; it was just no trouble. But maybe you could tell me, was it actually harder?
Well, there’s a couple things you are glancing over. Just a couple. And I suffer from the same problem, where you’ll jump through hoops without realizing that they’re hoops, because you’re just used to doing this kind of stuff… But the Twitter analogy works really well; so yes, there’s a shared namespace - and before that, people had email addresses, so it’s not like we’d lacked identity, but it did sort of unify those, so you know where to find anybody by a particular name, where to find a project by a particular name. But another thing that Twitter does too is it has a set of norms around how you communicate and how you do things with DMs and add replies and stuff like, right?
[00:27:59.29] That’s a really good point, yeah.
Source control is certainly part of the contribution experience, but if GitHub was just Git, it wouldn’t be the hub… It wouldn’t be GitHub, right? There’s an extension of the language and the tools around collaboration that they also unified. In Subversion I can create a diff, but how I send that diff to you and how you communicate that it may or may not go in or out, how we might communicate about that review process, that is not a unified experience across projects in older open source the way that it is in GitHub, right?
That’s true, and that’s a really good point. I mean, it was never hard to find out. Usually you mail the diff to the mailing list and people review it there, right? But you had to find out the address, you had to go read the project’s contribution documentation, and maybe that didn’t exist or was not easy to find… And you’re right, on GitHub it’s ‘Submit a pull request’. You know what to do - fork the repository, make your branch, make your change, turn it into a pull request against the upstream, and now it’s being tracked just like an issue, and by the way, the issue tracking is also integrated, so now you don’t have to go searching for the project’s issue tracker.
Yeah, I mean that workflow itself may not be more discoverable than sending a diff to a mailing list, but once you do it, it’s the same everywhere. I think that’s the bigger shift.
No, in fact I think it’s less discoverable, in the sense that the actual… I mean, I’ve trained a lot of people in using Git; I go to a wonderful organization… In fact, I’m gonna do a shout out for them, ChiHackNight.org, the Chicago Hack Night, on Tuesday nights here. There are a lot of newcomers there who haven’t used Git or GitHub before, or they’ve heard of it and tried it out. So I’ve had to walk people through this process of creating a PR, making their own fork or repository, and people get so confused, like “Wait, I’m forking the repository… But what’s a branch? What’s a repository? Where does the PR live?” It’s conceptually actually not easy at all, but once they know it, they know it for every project on GitHub. And I think your point is very good, it’s not that it’s easier, it’s just that you only have to learn it once now.
I think there’s also something to be said for the friendliness of GitHub, even just visually, right? Twitter is again maybe a great analogy for that… It’s just prettier. People feel more comfortable on a more consumer-facing website than navigating around the corners of the internet.
Yeah, and that’s one thing that Subversion never had - a default visual web browser interface. There were several of them and your project had to pick, so the one you picked might be different from what some other project picked. With GitHub it’s like… There are a lot of people who think of Git as GitHub. They think that that web interface that you see on GitHub, that is part of Git. Obviously, in some technical sense that’s not correct, but in a larger sense, as far as their experience and their actual workflow is concerned, that’s a pretty accurate way of looking at it.
Yeah. I think also - and this is one that is really easy to glance over if you have any experience, but because we’re in this new, publish-first mindset, newer people will publish stuff and put it up there, and they’ll actually get contributions. And it actually takes a much broader skillset to take contributions than it takes to push them to other projects, especially in traditional tooling, and GitHub also makes that incredibly easy. Their diff view is quite nice. They have the image different…
Yeah, it really is.
… and all of these other features, right? So if you’re somebody that doesn’t know Git very well and you just got your project up, getting a contribution and then having to pull it down locally and look the diff, it’s actually like a whole big extension of that collaboration toolchain, and they make that so easy for first-time publishers that are now dealing with contributions coming in. It makes that workflow for them really easy and it also just allows them to enjoy the process of getting contributions from people.
[00:32:00.07] Yeah, you’re right. I’ve never thought about that, but the process of becoming an open source maintainer is a lot easier on GitHub, and it’s so satisfying when you click that Merge Pull Request button and it just goes in. All you did was you clicked the green button and you’ve accepted a contribution from a perfect stranger on the internet. It’s so empowering, right? And that was not an easy process for new maintainers. In the old system you’d manually apply a diff and then commit it, and you’d have to write their name by hand in the log message, or something.
I think we’re also skipping over this entire generation of tools like Trac and JIRA, that in a lot of ways were much harder to use than sending a diff to a mailing list. [laughs]
Well yeah, I don’t know, because I got so used to them. I don’t think that they were a discrete generation; I think that they were a continuum of tools that as soon as the web came around, people started making bug trackers that… The original bug trackers were worked by email submission. You would communicate with them by sending them email and getting responses back, and actually a lot of projects ran on that. Then people started making websites that would track bugs and you could just interact with the website directly, and then that was integrated with Wiki functionality, Wiki was invented, and it just took a while for interfaces to sort out the conventions that actually worked. In a lot of ways, GitHub is the beneficiary of all the mistakes that everyone made before GitHub was created. If GitHub had been invented in the year 2000, they would have made all those same series of mistakes themselves, but instead they could just look back and see what everyone else did and not make those mistakes. No libel on them, of course, that’s what they should do, but that’s why it worked out so well for them.
It’s like MySpace and Facebook, or any sort of second adopter.
Yeah, exactly.
Well, I do think there’s another element of this though, which is that those tools - and JIRA in particular is very good at this… It’s developed for maintainers and for teams with a big project and a big process. So it is customizable to a project’s process. That means that’s great for that individual project if it exists alone by itself, but in an open source ecosystem where everytime I go to a JIRA there’s a different workflow, that’s incredibly daunting for individuals out there.
GitHub, because they were thinking about Git in the scale of people and contributions and forks and repos - you kind of take for granted that no, you can’t have super customized workflows at the repository.
Yeah… One of the things I kind of admire about GitHub’s management team is… I mean, if you look, GitHub has its own bug tracker. They have an open source code, but you can file bugs against GitHub itself, and that tracker is public. If you look through there, there are like thousands of these feature requests and modifications that people want, that for each person requesting, this change would suit their needs, it would really make life easier for their project, and basically GitHub employees spend their lives saying no. You just look in those threads and they are polite and they explain why, but they have to turn down most of those requests because they have to really think about the big picture and keep GitHub simple for the majority of open source projects, and they do a really good job at that.
[00:35:44.27] One of the things that I hope is happening, and I assume it is and I would like to look into it more is that GitLab and other open sourcers - in GitLab’s case there is an open source edition and also a proprietary edition - should be using GitHub as kind of like their free of charge research lab. All the things that are being requested in GitHub and all the decisions GitHub is making, and all the innovations that GitHub has to not experiment with because of their scale and all the existing customer base that they can’t afford to tick off - that is a real opportunity for these other platforms to say, “Hey, GitHub made the wrong call there. We’re gonna do that and try it out,” because they have less to lose right now and a lot to gain, and I think that there could be a very productive interplay between the two, that is in the long run good for open source. We’ll just have to see. But the fact that GitHub is making all these decisions in public is very useful, I think.
Yeah, I agree. So when you first got involved in open source in the ‘90s, it was sort of a counter-culture movement, and of all the things that you could say about open source today, I don’t think that you could say that it was a counter-culture movement.
Well, it’s funny… I think open source no longer thinks of itself as a counter-culture movement, especially in the United States. Well, actually let me back up a bit. So the term open source, at least for this usage of it, was coined in ’97, I think.
Right, right.
And open source was going on for many years prior to that. I had run an open source company and had been a full-time open source developer long before the term was coined, and people just used the term ‘free software’ and got confused, because there was just widespread confusion about whether that meant free as in there’s no charge. AOL used to ship CDs to everyone’s doorstep and that software was free, but it wasn’t free in the sense of free software, in the sense of freedom. So there was a lot of terminological confusion.
One of the things that I think is downplayed today, or there’s a little bit of historical amnesia about is the degree to which the coining of the term ‘open source’ was not simply an attempt to separate a development methodology from the ideological drives of the free software foundation Richard Stallman, but was also just an attempt to resolve a real terminology problem that a lot of people - and especially people who ran open source businesses - were having, which was “What term do we use that won’t confuse our customers and the people who use our software?”
Cygnus Solutions, which later got bought by Red Hat, tried to go with the term ‘sourceware’ for a while. That was an interesting coinage, and in fact my company, Cyclic Software, which I was running with Jim Blandy at the time, we actually contacted them to see about using that term, and we got a non-committal response where it wasn’t quite clear if they were trying to trademark it or they intended for only Cygnus to use it.
That’s even weirder.
So that didn’t fly, right? That wasn’t gonna work… If only Cygnus can use it, that’s not gonna be the term that kicks over.
That defeats the purpose, yeah.
Anyway, it didn’t have a good adjectival form, so it wasn’t… On its own merits, it had problems anyway. Eventually, when the term ‘open source’ came out, I just felt this tremendous relief. I was like, “Okay, no term is perfect. This term has some possible confusions and problems as well, but it is way easier for explanatory purposes than free software has been, so I’m just gonna start using it.” And I didn’t intend any ideological switch by that. I was still very pro free software, I ran only free software on my boxes, I only developed free software… But I just thought, “Okay, here’s a term that also means freedom that will confuse people less.”
[00:39:48.17] And then roughly a year after that coinage, when Stallman and the FSF (Free Software Foundation) realized that a lot of the people who were driving the term open source, who had founded the term - not necessarily people who were using the term, which was a lot of us - were also not on board with the ideology. Did they start to make this distinction between free software and open source, and say “Just because you support one doesn’t mean you support the other. They’re not the same thing, even though it’s the exact same set of licenses and software… So what do we mean by ‘not the same thing’?”
So that ideological split is kind of a post-facto creation. It was not actually something that was going on to the degree that it was later alleged to be going on.
And in your book, I’m trying to remember - it’s called Producing Open Source Software, but isn’t the subtitle also How To Run A Free Software Project?
Yeah, the book is a total diplomatic ‘split the difference’.
Yeah, you really went right down the middle there.
…How To Run A Successful Free Software Project. [laughs]
Yeah… You didn’t commit to either one.
Well, I didn’t want to, because to me it’s the same - like if there were two words for the vegetable broccoli, I might use both words, but it’s the same vegetable. Open source to me is one things; I can call it ‘free software’, I can call it ‘broccoli’, I can call it ‘open source’, it is still the same thing. People have all sorts of different motivations for doing it. Someone’s motivation for participating in a project or launching a project are not part of the project’s license, and therefore they’re not part of the term for me.
That’s a good transition into our next section. We’re gonna take a short break and when we come back we’ll talk about the mainstream version of open source.
[00:43:57.03] We’re back with Karl Fogel. Karl, today a lot of people are saying that open source is basically one, in the sense that a lot of companies are using it, a lot of people are roaring around the term ‘open source’ who might not have traditionally been engaged with open source… Do you think that open source has won, or they’re just sort of like different battles to be fought? Is that helpful vocabulary?
It has absolutely not won. I do not know why people think that. Where do you walk into a store and buy a mobile phone that’s running a truly open source operating system? I mean yeah, Android Core is open source, or is derived from the Android open source project. I guess when people say it’s won, what they mean is that if you think of software as a sphere where it’s constantly expanding - or as Marc Andreessen said “eating the world” - the surface of that sphere is mostly proprietary.
The ratio of the volume to the surface is constantly increasing, and most of that volume is open source, so people who are exposed to the backend of software and who are aware of what’s going on behind the scene in tech say, “Oh look, open source is winning” or “Open source has won” because so much of the volume inside the sphere is open source. But most of the world only has contact with the surface, and most of that surface is proprietary, and that surface is the only link that they’re going to have with any kind of meaningful software freedom, or lack of software freedom; their ability to customize, their ability to learn from the devices that they use… Their ability - I mean, it’s not the case that every person should be a programmer, but perhaps they should have the ability to hire someone else or bring something to a third party service that specializes in customization and get something fixed or made to behave in a different way. And for most of the surface of that sphere it’s completely impenetrable and opaque and you just can’t do that stuff; you have to accept what is handed to you. So no, I don’t think open source has won in the meaningful ways.
I think there’s a really important distinction there between software as infrastructure and software on the consumer-facing side. The research I’ve been doing and where I’m interested is almost exclusively on infrastructure, and I noticed there is this difference on maybe the ideals of free software to begin with, or around being able to change the Xerox printer, that was the Richard Stallman thing.
Right, that’s the legendary story, which I think is true, of Stallman trying to fix a printer and not having source code to the printer driver.
Right. And so I wonder, is that frustrating for them…? In some ways it really won on maybe the infrastructure side, and it’s almost even - I keep saying “won”, or just been massively adopted almost because it’s equivalent of free labour, like price-free stuff that startups can use, and so has the needle moved at all on the principle side of things? Or does it even matter?
Well, I have a very utilitarian view of the principle side of it; I do think that software freedom is important, but it’s increasingly an issue of control over your personal life and your families and friend’s lives, or at least being able not to put them in harm’s way. A great example is Karen Sandler, the executive director of the Software Freedom Conservancy, she has a heart device; she has a congenital heart condition, she has a device attached to her heart, and that device is running proprietary software. That software - I don’t know the exact version running on her device, but that type of software has been shown to be extremely vulnerable to hacking, to remote control.
[00:48:03.02] In fact Dick Cheney, the Vice-President had a similar device in his heart and apparently had the wireless features on the device disabled for security reasons. Think about the fact that the Federal Agency in the U.S. that is responsible for approving medical devices not only does not review software source code, it does not even require that the source code be placed in escrow with the Agency in case an investigation is later necessary. It just evaluates the entire system as a black box and says, “Yes, approved” or “No, not approved”, and they have nowhere near the resources or the confidence, let alone the mandate to review the software for vulnerabilities, when software vulnerabilities are increasingly affecting everyone. Everyone’s had a credit card account that’s been hacked in some way.
I wonder if those battles are gonna be addressed maybe not through software freedom or open source or those types of movements, but I guess as you’re describing it, I’m thinking more around hacker/maker movements and hardware stuff, or they might come at it from the same angle, saying “Why can’t I just modify anything?”
Yeah, and you do see a lot of that. I saw a keynote at the O’Reilly OSCon, the Open Source Convention, you probably saw it, too… The woman who had hacked her own insulin pump; the software that controls a device that dispenses a chemical into her bloodstream turned out to be hackable, so they hacked it.
So I think you’re right, the maker movement is driving it, and they share a lot of language and people with the open source movement. I just used the open source movement unironically; to me it’s largely the same as the free software movement.
So yeah, there are various pressures toward people having the ability to customize or to invite other people to help them customize the devices that run increasingly large swaths of our lives. I guess what’s happened is open source kept winning individual battles, but the number of things that software took a controlling role in kept increasing so rapidly that the percentage of things that are open source on the surface has been going down, even as open source keeps winning area after area.
I think that if you separated it nicely into two camps, if you look at the production of software versus the consumption of software, the reason we keep talking about “open source is winning” is because it really has won or very close to winning the production of software. If you were a developer in the early ‘90s, most if not all of your toolchain was proprietary. The way that you developed software was to use other proprietary software; that’s completely turned on its head.
Yeah, that was probably true, although it didn’t have to be.
It didn’t have to be at the time, but now the predominant way that you develop any software, including proprietary software, is to use a bunch of open source software.
Right, that’s a really good point. I think you’re right.
I mean, that proprietary code that’s on that hard device is probably compiled with JCC. [laughs]
Or one of the other free compilers.
Or LLVM, yeah. And so because the voices in our world are so dominated by the people that have actually produced the software, there is this mindset that “Hey, I live in this world all day that it’s 99% open source.” It feels like it has won. And I think the reason that it won though - in that space, and not in the consumer space - is that there is a utilitarian reason that you need something open source. It is infinitely more useful if it’s open source, and more useable as a producer if it’s open source. And there’s all these network effects that make it better over time that I can evaluate as a producer.
[00:52:12.24] But if you’re looking at products and the consumption of software, it being open source or not is not visible to the consumer of that software, at least not immediately. So there needs to be some kind of utilitarian argument around that, and I think it may be privacy and security. That’s a very, very good argument and it’s getting more tangible to consumers now.
Yeah, I think that’s at least part of it, and that has been a winning argument. A lot of the open source privacy and security projects have seen a lot more adoption and a lot more funding; just for various reasons, many of those projects tend to be non-profit, or at least not plausibly for-profit. It’s very clear that for all of his eloquence as a writer and speaker, which I think is considerable, the reason Richard Stallman succeeded was Emacs and GCC. He wrote or caused people to coalesce and help him write two really great programs, and then motivated a lot of people to write a lot of the pieces of the rest of a Unix-like system; didn’t unfortunately get the kernel, Linus Torvalds got that, and that has caused some bad blood ever since. But it was writing good code, that people could actually use, that gave him influence.
That’s why they took his other writings seriously, it was the utility of the code. But I think going back to the way you started presenting that idea, I think one of the important goals, one of the important motivating factors in the free software movement was keeping blurry the distinction between producers and consumers; the idea that there should not be a firm wall between these two camps, and that anyone who just thinks of themselves as only using software… I sort of prefer ‘user’ to ‘consumer’ because when you use software, you don’t - it’s not like apples, where once you use it, it’s consumed. [laughter] The software is still there after you run it, so it’s not being consumed. But the idea that any user has the potential, by very incremental degrees, to be invited into the production of the software… In fact, that’s what happened to me, that’s how I got into it. I was just using this stuff for writing my papers in college and exploring the nascent internet, and someone showed me how to use the info tree.
That was like the documentation tree for documentation that covered all of the GNU free software utilities, and right at the top of the introductory node, the top-level node in the info documentation browser was a paragraph that said “You can contribute to this documentation. To learn how to add more material to this info tree, click here”, where ‘click’ meant navigating with the keyboard and hit return; I don’t think there was a mouse. There was no mouse on those terminals, they were VT-100 terminals, but the idea that the system was inviting me to participate in the creation of more of the system - that struck me as really interesting.
[00:55:38.29] The idea was to keep the surface porous and allow for the possibility that those users who have the potential to become participants in improving the system do so. It wasn’t just freedom as an abstract concept, it was freedom as a practice. And still today, I think the way a lot of people get at open source is that they learn that they can affect the way a web page behaves by going behind the scenes and editing the JavaScript that got downloaded to their browser and noticing that things change; then they realize that “Hey, this is not a read-only system. The whole things is read/write. I can make things happen.” That’s what worries me about a lot of the user-facing devices and interfaces that we see today - there’s no doorway, there’s no porousness to the surface; you have no opportunity to customize or hack on it, or get in touch with the people who are one level closer to the source code.
I think there are a couple of interesting things that might be happening in tandem around that now. We haven’t talked about this at all, but just the definition of a software developer has changed radically in the past five years, where a lot more people are learning how to code. Maybe they’re not at a very high technical level, but just enough that they are able to modify small things around them and see that power. I think learning how to code has just become so much more accessible, so you have so many people that are interested in modifying the world around them in much more casual ways. That is blurring the line between consumer and producer. Look at any child today, everybody is learning how to code, and just imagine when they grow up and they just expect that everything around them can be transformed. It’s almost like people are coming at it from a different direction, but then at the same time you see all these very proprietary platforms that are basically exploiting network effects to centralize where people congregate on the internet, and those things are still total black boxes.
I don’t know what happens when the youngest generation now grows up… Will they say, “This is bullshit!”? “This is not how we were raised to see the internet.”
They’ll say “This is bullshit”, but they’ll say it on Facebook.
Right! And that’s the hard part, is sort of like you have this tyranny of … yeah.
I think that point about network effects is really important. What happened as an increasingly large percentage of humanity got internet connections was that the payoff ratio for building a proprietary system changed. It used to be that if you were building a system there was some reward for making it a little hackable, because the users you were likely to attract… Well, people on the internet at that time were already more likely to have potential to contribute to your system, so there was statistically some potential reward for making your system have a slightly open door to people coming in and helping out. But if you’re launching something like Facebook or Snapchat in the age of most of humanity being online, then the trouble you go through to make that thing hackable versus the payoff when most of those users are not going to take advantage of that, the reward matrix just looks different now, and maybe it just doesn’t make economic sense for those proprietary platforms to have a porous surface.
And oddly you see, like on Snapchat for example, where people are… Snapchat offers tons of things to make people essentially modify around them, like stickers, drawing on things or whatever. So it’s that same behavior, but it’s still on Snapchat’s platform.
Right, and they control it and they track… Like, you can’t fork Snapchat and make your stickers in the forked Snapchat, let alone do something else.
[00:59:41.22] The uncharitable way to say it is that everyone’s creative and environmental improvement impulses are being coopted and redirected into limited and controlled actions that do not threaten the platform providers. Basically every platform provider’s business model is “I wanna be like a phone carrier. I just wanna have total control over the user base and have people have to join in order to get access to the rest of my user base”, and that creates a mentality that is antithetical to the way open source works. You don’t fork a monopoly-based thing. You don’t fork a thing that has network effects.
I have a hard time thinking that that is necessarily… That these things have to be in conflict. I don’t think that users are ever gonna… I don’t think that you can sell a product to users in a competitive market based on the values that will attract a community around people hacking it. You have to be a great product compared to everybody else on the terms that most users are using it, but that doesn’t necessarily mean that you can’t also be hackable. You just have to have a culture around the product of actually creating something good.
Look at the one success story that we have, for a short period of time, which was Mozilla. They won for a while and took a huge amount of market share away from Microsoft - enough that Microsoft actually came and participated at Web Standards again - because they made a better browser for users, and not just for people that were hacking on websites.
And it’s because it’s better, not necessarily because of those…
Oh no, my doom and gloom is not a moral condemnation, it’s an observation of economic reality. I think what you’re saying is correct, but it’s still not good news for open source.
No, and I think that’s what’s so interesting about right now in even how people are using the term open source, and a lot of people say something is open source when it’s not actually. So the term itself has been sort of being coopted into different definitions, and for a lot of people now that are just coming into it, they say the term ‘open source’ and they just mean “Why shouldn’t I share what I made with the world?” or “Why shouldn’t I change something that I see?”, but it doesn’t necessarily carry all that other history or expectations with it.
Well yeah, that coopting has been going on. Ever since the term was coined, there have been groups and people using it in ways that don’t mean what it originally meant. There have been people coopting the term since it was coined, but there’s always been counter pressure to preserve its original meaning, because the original meaning is so unambiguous and so clear. It’s so easy to identify when it’s being correctly used, that the counter pressure usually is successful. So I don’t see any more of that now than in the past. I think that’s just a constant terminological tug of war that’s going on, but mostly the meaning of the term is as strong now as it ever was.
Well, I think it’s as strong now to a set of people that still hold on to that term really strongly, but to be frank I think they’re almost putting blinders on to how so many other people are using it. We’ve talked about this - at what point does that new definition just become the definition because so many people are using it that way?
Yeah, that’s how the language works and I’m totally on board with that, but I guess what I’m saying is I try to see that happening - and a number of people do, and then they actually go where possible… When it’s an organizational source of terminology dilution, they’ll go to that organization and say, “Hey, the term doesn’t mean that. Stop doing that!” and in almost every case the organization reforms their usage, and that’s the only reason that open source still means anything; it’s because that constant process is going on, and I haven’t actually seen the ratio changing that much lately, and of course it’s a very hard thing to gather data on, and Nadia you have been trying to gather data on this and you’ve been out there doing research on this so you might be right, but the blinders are anyway not intentional. We are actually out actively looking for that, and to me it looks like it’s about the same as it ever was, and we just have to stay vigilant.
[01:04:08.23] That’s a nice recap of the problems of people misusing the term or using it for something that’s not within the scope of what open source means. But there’s also a fair amount of - I don’t know how to say this without being mean…
Oh, go for it.
Corporations or projects that are open source within the definition of open source, but aren’t what we would call open.
Actually, I think that’s okay and I don’t care. In other words, if you’re forkable, you’re open source. And if you run the project as a closed society and even the developer’s names are kept top secret, as long as the source code is available and its under an open source license and it could be forked, you’re open source.
You’re thinking more about the future of it, rather than the current reality. Like, even if I can’t get anything done now, if it becomes a big enough problem, I have that option, right?
Well yeah. I mean, the fact that you have that option affects the behavior of the central maintainers, whether they admit it openly or not. The knowledge that your thing can be forked causes you to maintain it differently, even if you never respond to any of the pull requests, you never respond to any of the emails of anyone from outside the maintainer group. The mere fact that someone could fork it forces you to take certain decisions in certain directions so as not to increase the danger of forking, for example. So you still get open source dynamics, even when they’re not visible.
Yeah, that’s a good question, Nadia. I do think that some people put blinders on and try to ignore it, but they tend to get reminded of it. [laughs]
I didn’t hear Nadia’s question, I’m sorry.
I really wonder whether some companies actually see it that way, or whether they’re actually acutely aware of the fear of a fork. Because again, like we talked about network effects, where even if nobody likes the thing anymore, if everybody is using a certain thing, it’s very hard to actually switch off.
Well, it just requires… I mean, for business-to-business open source. Again, Android is a classic example. Google is very aware of the potential for forks; they are very aware of the business implications, to the extent that those are predictable, depending on who might fork it. And indeed, some forks have started to appear, and that is something that gets factored into their decisions as to how they run their copy of the Android project, which so far most companies still socially accept as the master copy, but they are not required to do that. So that means at least the Android Core code is indeed open source, even though it is not run in the way most open source projects are; although I think actually they have taken contributions from the outside. It’s not quite as closed as the tech press indicates it is.
From what I understand of your views, you see it as like the license and these guaranteed freedoms are what makes it open source and that’s all it really matters, because you’re saying if need you could always fork it.
I’m not quite saying that that’s all that really matters, I’m just saying that it’s a main thing… And sure, I would much rather have a project be run by a community, but that potential is always there as long as the open source license is there.
Yeah, the reason why I think collaboration and community is so intertwined is because, again, network effects… And it doesn’t really matter whether something can technically be forked if there is actually no ability to change it, so I worry that relying too much on that core definition could act… It’s sort of like this great hypothetical about whether that really happens. It’s like anyone can create an alternative to Facebook in theory, but no one has successfully created an alternative, because everyone’s on Facebook.
[01:08:04.11] Well, but I don’t think that network effects in an open source development environment are quite the same… Let’s take a couple of examples. GCC got forked years ago. It had a core group of maintainers, and then it had a bunch of revolutionaries who were not happy with how those maintainers were maintaining it. And from the beginning of the project there was no doubt about who this sort of socially accepted master copy was. It was the one maintained by the Free Software Foundation with a technical council that I don’t know how they were selected, but I think Richard Stallman was involved in selecting them, and when these revolutionaries grew increasingly unhappy with technical decision being made and with how contributions were being accepted or not accepted, they had corporate funding, they went off and created EGCS.
EGCS started accepting all those patches that the GCC copy wouldn’t take, and eventually it kind of blew past GCC in terms of technical ability to the point where the FSF said “Well, I guess you’re kind of where stuff is happening now, so we’re just gonna take the next version of EGCS and call that GCC and merge the two, and you won.” And it was totally successful, and it happened because the problems were big enough that people were willing to devote resources to forking and solving them. Could the same thing happen with the Linux kernel? Absolutely. If Linus started making bad decisions, or if he started ticking off too many people and enough kernel developers who had the technical plausibility to launch a fork chose to make a fork - yeah, it would succeed, there’s no question. But it’s just that Linus is running the project well enough that no one needs to do that.
Yeah, I see your point, it is different.
Yeah, but Facebook, on the other hand, that’s a whole different kind of network effect. I don’t mean to completely argue your point away because I think it’s a good one, which is that there are network effects, and it is a lot of effort to fork a popular project that has a successful or at least a cohesive maintenance team and a clear leadership structure.
And you need to have a community that cares enough to fork it. Again, fast-forwarding to some sort dystopian future that I don’t actually know is the future or not, but if open source projects become more about users than about contributors, and people are just sort of using the thing, then it becomes a lot harder to mobilize people to change something. But maybe I’m just sort of making up…
Well, the degree… The ease with which it is possible to motivate people to make a fork or to change something will always be directly proportional to the amount of need for that change. If no one’s motivated to change anything, that just means it’s not important to someone for something to get changed, so why should we care?
Yeah, I don’t know if–People can hate using something… There’s a ton of legacy open source projects that are used in everybody’s code and it’s just really hard to switch out because everyone uses them.
I think the difference though is that there’s just not enough people… Yes, people hate using it, but there’s not enough people that want to be developing on it that can’t, that would then fork it and fix it. And think that there’s a tension here between the people using it and the people that wanna contribute and can’t, or wanna fix this and can’t. And sometimes it really is too difficult to pull that out. But io.js was a pretty successful fork, and that was in large part because there were a lot of people that wanted to contribute that couldn’t, and that wanted to take on ownership of the project and couldn’t. So there was a thriving community actually working on it, and then people that were using were like “Oh great, I can come and use this.”
[01:12:00.15] Unfortunately I don’t know the details of that particular fork, it sounds like you do. If you think there are interesting lessons to draw from it, please explain more.
So I’ve said this on a couple occasions, but I think the size of the user base is proportional… There’s some percentage of that that would contribute, that wanna contribute in some way, and if they’re enabled to, you’ll have a thriving community. If you don’t, you eventually will increase the tension, not just with your overall user base, but also with these people that would be contributing. And eventually, if that tension rises enough, you get a fork.
I think that where that starts to pare down is that when you look at Android, the users of the Android code base are not the users of Android. The users of the Android code base are companies that manufacture phones, for the most part.
And indeed, they started forking Android.
Yes, exactly. So they have the resources to do that, and their needs do not necessarily line up with the needs of Google. The problem is that their needs are in many cases counter to the users of Android, so it puts Google in a strange place where they’re not satisfying the needs of the users of the Android code base, but they are satisfying the needs of the Android end users. If you talk to anybody who uses Android, they’re like “Oh, I have to use the newest Google phone that only takes the Google Android, because the ones where manufacturers have forked them are pretty much terrible.” Except, I heard Java is really good. I think we’re getting into very specific things right now… [laughter]
Well we are, but just to make a quick point about that, in theory, in some sort of long ark of software justice, there should be a link between what those companies are doing with their forks of Android and user’s needs, because otherwise they’re not gonna sell phones. Of course, I would love all those phones to be running a fully open source operating system, and the reasons why they’re not are an interesting topic in their own right, but there should be some connection eventually between those forks and some kind of technical need being solved.
So when you’re looking towards the future though, do you see that tension rising, and users starting to come more in conflict with that model, or are you more pessimistic about it and you feel like the surface is going to continue to be dominated the way that it is now?
I wanna give the optimistic answer, but I have no justification for it. Because software is increasingly being tied to hardware devices, and the hackability for a hardware device is so much… Like, the hacktivation energy, the threshold for hacking on something other than a normal laptop or desktop computer is just so much higher that the ratio in any given pool, in any given user base, the number of those users who will be developers, the percentage is gonna be lower. Just to hack on an Android phone - alright, you’ve gotta setup an Android development environment, you’ve gotta plug into the phone using a special piece of software that gets you into a development environment, and all of that software might be open source, but it’s not like just compiling and running a program and then hacking on the source code and running it again on your laptop. The overhead to get to the point of development is just so much higher. And that’s just phones. Do you think hacking on your car is gonna be easier than that? No, it’s gonna be a lot harder.
I think unfortunately we have to leave it there with this view of a dystopian future… [laughter]
Always happy to make it darker for you.
…but we’ll be back next week. We’re gonna continue with Karl and talk about some much happier things, like contributions and governance models…
Oh, I’ll turn that dark, too.
Oh, okay. [laughter]
Can’t wait!
Our transcripts are open source on GitHub. Improvements are welcome. 💚 | https://changelog.com/rfc/1 | CC-MAIN-2020-16 | en | refinedweb |
Inheritance in c4d [SOLVED]
On 13/06/2015 at 04:10, xxxxxxxx wrote:
Hi!
I was having some doubt on inheritance map in module c4d in c4d python sdk.It would be of gr8 help if you help me out. Here it is:
As I understand c4d is a module name. from
link it seems that all other boxes on that page are just classes. right? Now, if i look closely, say c4d.plugins.BaseData, it is parent of lots of , say , c4d.plugins.xyz. Now the thing I am confusing is that : we use c4d[dot] becoz we are using "import c4d" in the header. Why do we have plugin[dot] when there is no plugin module. as in is "plugin.xyz" literally a name of a class or is plugin a module or plugin a function??
On 14/06/2015 at 12:41, xxxxxxxx wrote:
Hi @CreamyGiraffe,
"import c4d" imports the "c4d" package. Some of the subpackages only include c4d classes, however, some of the submodules (forgive me, my understanding of package vs module, etc is a little fuzzy) just contain functions, not actual classes. Some, like "plugins" contain both classes and functions directly accessible under the "c4d.plugins" namespace.
"c4d.plugins" gives you access to all of the Classes and functions inside of c4d.plugins. "c4d.plugins.BasePlugin()" lets you create an object of type BasePlugin, whereas "c4d.plugins.GeLoadString()" calls the function GeLoadString.
You'll have to consult the documentation to see what is a subclass and what is just a function.
Cheers,
Donovan
On 14/06/2015 at 23:44, xxxxxxxx wrote:
Thanks Donovan | https://plugincafe.maxon.net/topic/8822/11657_inheritance-in-c4d-solved | CC-MAIN-2020-16 | en | refinedweb |
How do I use a progress bar when my script is doing some task that is likely to take time?
For example, a function which takes some time to complete and returns
True when done. How can I display a progress bar during the time the function is being executed?
Note that I need this to be in real time, so I can't figure out what to do about it. Do I need a
thread for this? I have no idea.
Right now I am not printing anything while the function is being executed, however a progress bar would be nice. Also I am more interested in how this can be done from a code point of view.
There are specific libraries (like this one here) but maybe something very simple would do:
import time import sys toolbar_width = 40 # setup toolbar sys.stdout.write("[%s]" % (" " * toolbar_width)) sys.stdout.flush() sys.stdout.write("\b" * (toolbar_width+1)) # return to start of line, after '[' for i in xrange(toolbar_width): time.sleep(0.1) # do real work here # update the bar sys.stdout.write("-") sys.stdout.flush() sys.stdout.write("]\n") # this ends the progress bar
Note: progressbar2 is a fork of progressbar which hasn't been maintained in years.
With tqdm you can add a progress meter to your loops in a second:
In [1]: import time In [2]: from tqdm import tqdm In [3]: for i in tqdm(range(10)): ....: time.sleep(3) 60%|██████ | 6/10 [00:18<00:12, 0.33 it/s]
Also, there is a graphical version of tqdm since
v2.0.0 (
d977a0c):
In [1]: import time In [2]: from tqdm import tqdm_gui In [3]: for i in tqdm_gui(range(100)): ....: time.sleep(3)
But be careful, since
tqdm_gui can raise a
TqdmExperimentalWarning: GUI is experimental/alpha, you can ignore it by using
warnings.simplefilter("ignore"), but it will ignore all warnings in your code after that. | https://pythonpedia.com/en/knowledge-base/3160699/python-progress-bar | CC-MAIN-2020-16 | en | refinedweb |
This is a guest post co-written by Solarflare, a Xilinx company. Miklos Reiter is Software Development Manager at Solarflare and leads the development of Solarflare’s Cloud Onload Operator. Zvonko Kaiser is Team Lead at Red Hat and leads the development of the Node Feature Discovery operator.
Figure 1: Demo of Onload accelerating Netperf in a Pod
Solarflare, now part of Xilinx, and Red Hat have collaborated to bring Solarflare’s Cloud Onload for Kubernetes to Red Hat’s OpenShift Container Platform. Solarflare’s Cloud Onload accelerates and scales network-intensive applications such as in-memory databases, software load balancers and web servers. The OpenShift Container Platform empowers developers to innovate and ship faster as the leading hybrid cloud, enterprise Kubernetes container platform.
The Solarflare Cloud Onload Operator automates the deployment of Cloud Onload for Red Hat OpenShift and Kubernetes. Two distinct use cases are supported:
- Acceleration of workloads using Multus and macvlan/ipvlan
- Acceleration of workloads over a Calico network
This blog post describes the first use case; a future blog post will focus on the second use case.
Solarflare's Cloud Onload Operator provides an integration path with Red Hat OpenShift Container Platform's Device Plugin framework, which allows OpenShift to allocate and schedule containers according to the availability of specialized hardware resources. The Cloud Onload Operator uses the Multus multi-networking support and is compatible with both the immutable Red Hat CoreOS operating system as well as Red Hat Enterprise Linux. The Node Feature Discovery operator is also a part of this story, as it helps to automatically discover and use compute nodes with high-performance Solarflare network adapters, which Multus makes available to containers in addition to the usual Kubernetes network interface. OpenShift 4.2 will include the Node Feature Discovery operator.
Below is a network benchmark showing the benefits of Cloud Onload on OpenShift.
Up to 15x Performance Increase
Figure 2: NetPerf request-response performance with Onload versus the kernel
Figure 2 above illustrates the dramatic acceleration in network performance, which can be achieved with Cloud Onload. With Cloud Onload, a NetPerf TCP request-response test produces a more significant number of transactions per second delivering better performance than with just the native kernel network stack.
Moreover, performance scales almost linearly as we scale the number of NetPerf test streams up to the number of CPU cores in each server. In this test, Cloud Onload achieves eight times the kernel transaction rate with one stream, rising to a factor of 15 for 36 concurrent streams.
This test used a pair of machines with Solarflare XtremeScale X2541 100G adapters connected back-to-back (without going via a switch). The servers were using 2 x Intel Xeon E5-2697 v4 CPUs running at 2.30GHz.
Integration with Red Hat OpenShift
Deployment of Onload Drivers
The Cloud Onload Operator automates the deployment of the Onload kernel drivers and userspace libraries in Kubernetes.
For portability across operating systems, the kernel drivers are distributed as a driver container image. The operator ships with versions built against Red Hat Enterprise Linux and Red Hat CoreOS kernels. For non-standard kernels, one can build a custom driver container image. The operator automatically runs the driver container on each Kubernetes node, which loads the kernel modules.
Also, the driver container installs the user-space libraries on the host. Using a device plugin, the operator then injects the user-space libraries, together with the necessary Onload device files, into every pod which requires Onload.
Deployment of Onload on Kubernetes is significantly simplified by the operator, as it is not necessary to build Onload into application container images or to write custom sidecar injector or other logic to achieve the same effect.
Configuring Multus
OpenShift 4 ships with the Multus multi-networking plugin. Multus enables the creation of multiple network interfaces for Kubernetes pods.
Before we can create accelerated pods, we need to define a Network Attachment Definition (NAD) in the Kubernetes API. This object specifies which of the node's interfaces to use for accelerated traffic, and also how to assign IP addresses to pod interfaces.
The Multus network configuration can vary from node to node, which is useful to assign static IPs to pods, or if the name of the Solarflare interface to use varies between nodes.
The following steps create a Multus network that can provide a macvlan subinterface for every pod that requests one. The plugin automatically allocates static IPs to configure the subinterface for each pod.
First, we create the
NetworkAttachmentDefinition(NAD) object:
cat << EOF | oc apply -f -
apiVersion: "k8s.cni.cncf.io/v1"
kind: NetworkAttachmentDefinition
metadata:
name: onload-network
EOF
Then on each node that uses this network, we write a Multus config file specifying the properties of this network:
mkdir -p /etc/cni/multus/net.d
cat << EOF > /etc/cni/multus/net.d/onload-network.conf
{
"cniVersion": "0.3.0",
"type": "macvlan",
"name": "onload-network",
"master": "sfc0",
"mode": "bridge",
"ipam": {
"type": "host-local",
"subnet": "172.20.0.0/16",
"rangeStart": "172.20.10.1",
"rangeEnd": "172.20.10.253",
"routes": [
{ "dst": "0.0.0.0/0" }
]
}
}
EOF
Here,
master specifies the name of the Solarflare interface on the node, and
rangeStart and
rangeEnd specify non-overlapping subsets of the subnet IP range.
An alternative to the macvlan plugin is the ipvlan plugin. The main difference between the ipvlan subinterfaces and the macvlan is that the ipvlan subinterfaces share the parent interface’s MAC address, providing better scalability in the L2 switching infrastructure. Cloud Onload 7.0 adds support for accelerating ipvlan subinterfaces in addition to macvlan subinterfaces. OpenShift 4.2 will add support for ipvlan.
Node Feature Discovery
A large cluster often runs on servers with different hardware. This means that workloads requiring high-performance networking may need to be explicitly scheduled to nodes with the appropriate hardware specification. In particular, Cloud Onload requires Solarflare XtremeScale X2 network adapters.
To assist with scheduling, we can use Node Feature Discovery. The NFD operator automatically detects hardware features and advertises them using node labels. We can use these node labels to restrict which nodes are used by the Cloud Onload Operator, by setting the Cloud Onload Operator’s
nodeSelector property.
In future, NFD will be available within the operator marketplace in OpenShift 4.2. At the time of writing, NFD is installed manually as follows:
$ git clone
$ cd cluster-nfd-operator
$ make deploy
We can check that NFD has started successfully by confirming that all pods in the openshift-nfd namespace are running:
<span>$ oc get pods -n openshift-nfd</span>
At this point, all compute nodes with Solarflare NICs should have a node label indicating the presence of a PCI device with the Solarflare vendor ID (0x1924). We can check this by querying for nodes with the relevant label:
<span>$ oc get nodes -l feature.node.kubernetes.io/pci-1924.present=true</span>
We can now use this node label in the Cloud Onload Operator’s
nodeSelector to restrict the nodes used with Onload. For maximum flexibility, we can, of course, use any node labels configured in the cluster.
Cloud Onload Installation
Installation requires downloading a zip file containing a number of YAML manifests from the Solarflare support website. Following the installation instructions in the README.txt contained in the zip file, we edit the example custom resource to specify the kernel version of the cluster worker nodes we are running:
<span>kernelVersion: "4.18.0-80.1.2.el8_0.x86_64”</span>
and the node selector:
nodeSelector:
beta.kubernetes.io/arch: amd64
node-role.kubernetes.io/worker: ''
feature.node.kubernetes.io/pci-1924.present: true
We then apply the manifests
<span>$ for yaml_spec in manifests/*; do oc apply -f $yaml_spec; done</span>
We expect to list the Solarflare Cloud Onload Operator on soon, for installation using the Operator Lifecycle Manager and OpenShift’s built-in Operator Hub support.
Running the NetPerf Benchmark
We are now ready to create pods that can run Onload.
Netperf Test Image
We now build a container image which includes the netperf performance benchmark tool using a Fedora base image. Most common distributions that use glibc are compatible with Onload. This excludes extremely lightweight images, such as Alpine Linux.
The following Dockerfile produces the required image.
NetPerf.Dockerfile:
FROM fedora:29
RUN dnf -y install gcc make net-tools httpd iproute iputils procps-ng kmod which
ADD /root
RUN tar -xzf /root/netperf-2.7.0.tar.gz
RUN netperf-netperf-2.7.0/configure --prefix=/usr
RUN make install
CMD ["/bin/bash"]
We build the image:
<span>$ docker build -t netperf -f netperf.Dockerfile</span>
Example onloaded netperf daemonset
This is an example daemonset that runs netperf test pods on all nodes that have Solarflare interfaces.
apiVersion: apps/v1
kind: DaemonSet
metadata:
name: netperf
spec:
selector:
matchLabels:
name: netperf
template:
metadata:
labels:
name: netperf
annotations:
k8s.v1.cni.cncf.io/networks: onload-network
spec:
nodeSelector:
node-role.kubernetes.io/worker: ''
containers:
- name: netperf
image: /netperf:latest
stdin: true
tty: true
resources:
limits:
solarflare.com/sfc: 1
Here
is the registry hostname (and :port if required).
The important sections are:
- The
annotationssection under
spec/template/metadataspecifies which Multus network to use. With this annotation, Multus will provision a macvlan interface for the pods.
- The
resourcessection under
containersrequests Onload acceleration from the Cloud Onload Operator.
Running Onload inside accelerated pods with OpenShift/Multus
Each netperf test pod we have created has two network interfaces.
eth0: the default Openshift interface
net1: the Solarflare macvlan interface to be used with Onload
Any traffic between the net1 interfaces of two pods can be accelerated using Onload by either:
- Prefixing the command with "onload"
- Running with the environment variable LD_PRELOAD=libonload.so
Note: One caveat to the above is that two accelerated pods can only communicate using Onload if they are running on different nodes. (Onload bypasses the kernel's macvlan driver to send traffic directly to the NIC, so traffic directed at another pod on the same node will not arrive.)
To run a simple netperf latency test we open a shell on each of two pods by running:
$ kubectl get pods
$ kubectl exec -it <pod_name> bash
On pod 1:
$ ifconfig net1
net1: flags=4163<UP,BROADCAST,RUNNING,MULTICAST> mtu 1500
inet 172.20.0.16 netmask 255.255.0.0 broadcast 0.0.0.0
[...]
bash-4.4$ onload --profile=latency netserver -p 4444
oo:netserver[107]: Using OpenOnload 201811 Copyright 2006-2018 Solarflare Communications, 2002-2005 Level 5 Networks [4]
Starting netserver with host 'IN(6)ADDR_ANY' port '4444' and family AF_UNSPEC
On pod 2:
<span>$ onload --profile=latency netperf -p 4444 -H 172.20.0.16 -t TCP_RR</span>
Running multiple parallel NetPerf pairs, concurrently produced the results shown above.
Obtaining Cloud Onload
Visit to learn more about Cloud Onload or make a sales inquiry. An evaluation of Solarflare’s Cloud Onload Operator for Kubernetes and OpenShift can be arranged on request.
Categories
Kubernetes, News, OpenShift Container Engine, OpenShift Container Platform, How-tos, Operator Framework, Containers | https://www.openshift.com/blog/launching-openshift-kubernetes-support-for-solarflare-cloud-onload | CC-MAIN-2020-16 | en | refinedweb |
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