text stringlengths 454 608k | url stringlengths 17 896 | dump stringclasses 91
values | source stringclasses 1
value | word_count int64 101 114k | flesch_reading_ease float64 50 104 |
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ok im in an entry level class and very new to java so your help would be greatly appreciated..here is what were supposed to do...
Write a program to take bids for an auction business. The program prompts the user to enter the name of item desired, next prompts the user to enter the bid price, next prompt the user to enter the desired method of shipping. There are three shipping methods available. ( 0 means ground shipping four day delivery, 1 means ground express two day delivery, 2 means air shipping one day delivery ). No charge for ground four day delivery, there is a $5.00 charge for ground express and $15.00 charge for air shipping ). Assume that the user of the program will be perfectly alert and a perfect typist, will not make any mental or typing mistakes.
A sample run would look like this :
Enter Item you are bidding on : bicycle
Enter your bid price : 50.00
Enter desired delivery method ( 0 for ground four day, 1 for ground express, 2 for air shipping ) : 1
Invoice :
Item : Bicycle
Your bid price : 50.00
shipping : 5.00
total : 55.00
here is my program so far...
import java.util.Scanner; public class Assign4_Roberts{ public static void main(String[] args){ //enter shipping prices double ground = 0; double express = 5; double air = 15; //create a scanner Scanner input = new Scanner(System.in); System.out.println("Enter the item you are bidding on :"); int answer =input.nextInt(); System.out.println("Enter your maximum bid: "); int num = input.nextInt(); System.out.println("Enter desired delivery method ( 0 for ground four day, 1 for ground express, 2 for air shipping ) :"); int shipping = input.nextInt(); if(shipping == 0) System.out.println("total:" + num + ground); if(shipping == 1) System.out.println("total:" + num + express); if(shipping == 2) System.out.println("total:" + num + air); //invoice System.out.println("Invoice :"); System.out.println("Item :" + answer); System.out.println("Your Bid Price :" + num); System.out.println("Total :" + num + shipping); } }[B][/B]
and here is what happens when i try to run the program...im using eclipse btw
Enter the item you are bidding on :
bicycle
Exception in thread "main" java.util.InputMismatchException
at java.util.Scanner.throwFor(Unknown Source)
at java.util.Scanner.next(Unknown Source)
at java.util.Scanner.nextInt(Unknown Source)
at java.util.Scanner.nextInt(Unknown Source)
at Assign4_Roberts.main(Assign4_Roberts.java:20)
anybody know what im doing wrong | http://www.javaprogrammingforums.com/whats-wrong-my-code/5435-need-help-homework-program-thanks.html | CC-MAIN-2016-36 | refinedweb | 405 | 52.76 |
I need to write a function that creates a list of tuples of the first occurrence of a letter followed by its row and column in a list of lists.
Example Input and Output:
#Input:
lot2 = [['.','M','M','H','H'],
['A','.','.','.','f'],
['B','C','D','.','f']]
#Output: [('M', 0, 1), ('H', 0, 3), ('f', 1, 4), ('B', 2, 0)]
letter = '.abcdefghijklmnopqrstuvwxyzABCDEFGHIJKLMNOPQRSTUVWXYZ'
def list_cars(lst):
for y, row in enumerate(lst):
if letter in row:
return letter, y, row.index(letter)
First off, use the string library to get a string of all upper and lower case letters:
import string string.ascii_letters Out[40]: 'abcdefghijklmnopqrstuvwxyzABCDEFGHIJKLMNOPQRSTUVWXYZ' collector = [] output_list = [] for i in lot2: for j in i: if j in string.ascii_letters and j not in collector: tmp = (j,lot2.index(i), i.index(j)) output_list.append(tmp) collector.append(j)
output_list should give you what you want.
edit: If you want to also capture full-stops use string.printable - although this will give you a string that consists of additional punctuation and white space characters as well. | https://codedump.io/share/aLK9sJwOvXlG/1/create-a-list-of-lists-of-tuples-where-each-tuple-is-the-first-occurrence-of-a-letter-along-with-its-row-and-column-in-the-list-of-lists | CC-MAIN-2018-26 | refinedweb | 175 | 64.71 |
OpenGL utility to render a vil_image_view. More...
#include <vgui_vil_image_renderer.h>
OpenGL utility to render a vil_image_view.
This is not a tableau.
The vil_image_resource (as given) The image is rendered by extracting appropriate views from the resource There are different modes of rendering based on flags set on the vgui_range_map_params range mapping class. The states are:
1) rmp == null - a buffer is created for the entire image and the buffer supplies glPixels to the display. Range map tables are used where appropriate. No pyramid image support
2) rmp->use_glPixelMap && rmp->cache_mapped_pix pyramid images are handled properly and the range maps are used to generate the scaled pyramid view buffer. gl hardware maps are not used. gl pixels are generated in software using the maps. Rendering is limited to the current viewport. The buffer is updated only if the viewport changes
3) rmp->use_glPixelMap && !rmp->cache_mapped_pix pyramid images are handled properly and the range maps are loaded into gl hardware for rendering. The pyramid image data that drives the hardware is cached as a memory_chunk, but not a pre-mapped glPixel buffer as in 2). The memory_chunk is only updated if the viewport changes.
Definition at line 59 of file vgui_vil_image_renderer.h.
Constructor - create an empty image renderer.
Definition at line 33 of file vgui_vil_image_renderer.cxx.
Destructor - delete image buffer.
Definition at line 41 of file vgui_vil_image_renderer.cxx.
Create a buffer if necessary.
Definition at line 79 of file vgui_vil_image_renderer.cxx.
Create a buffer with viewport dimensions.
creates a buffer for a portion of the image.
Definition at line 108 of file vgui_vil_image_renderer.cxx.
Create a buffer from specified resource corresponding to a pyramid zoom level.
Definition at line 122 of file vgui_vil_image_renderer.cxx.
draw the pixels to the frame buffer.
Definition at line 142 of file vgui_vil_image_renderer.cxx.
Return the image resource that this renderer draws.
Definition at line 63 of file vgui_vil_image_renderer.cxx.
Are the range params those used to form the current buffer.
Are the range map params associated with the current buffer out of date?.
If so we have to refresh the buffer.
Definition at line 469 of file vgui_vil_image_renderer.cxx.
Renders the image pixels. If mp not null then render over an interval.
Definition at line 445 of file vgui_vil_image_renderer.cxx.
Render the pixels in hardware using the glPixelMap with range_map data.
Note that some OpenGL environments have no graphics hardware but the glPixelMap is still somewhat faster JLM (on a DELL precision)
Definition at line 149 of file vgui_vil_image_renderer.cxx.
Tell the renderer that the underlying image data has been changed.
Definition at line 71 of file vgui_vil_image_renderer.cxx.
Attach the renderer to a new view.
Definition at line 48 of file vgui_vil_image_renderer.cxx.
Stored the GL pixels corresponding to the image data.
Definition at line 89 of file vgui_vil_image_renderer.h.
a cache for the range map params associated with buffer.
Definition at line 92 of file vgui_vil_image_renderer.h.
Stores the image data (pixels, dimensions, etc).
Definition at line 86 of file vgui_vil_image_renderer.h.
buffer state variable.
Definition at line 95 of file vgui_vil_image_renderer.h.
a cache when rendering using the gl hardware map.
Definition at line 98 of file vgui_vil_image_renderer.h. | http://public.kitware.com/vxl/doc/release/core/vgui/html/classvgui__vil__image__renderer.html | crawl-003 | refinedweb | 526 | 61.83 |
Eclipse Community Forums - RDF feed Eclipse Community Forums DLTK indexing <![CDATA[I am trying to use the Python SDK before implementing dltk in a separate language. But till now I am not able to make it work as a whole. The parsing seems to work correctly and even I am getting the outline of the code in the outline view. Now my next move is to verify the python search. But this is not working. Below is the PythonSearchFactory class. ]public class PythonSearchFactory extends AbstractSearchFactory { public IMatchLocatorParser createMatchParser(MatchLocator locator) { return new PythonMatchLocationParser(locator); } public SourceIndexerRequestor createSourceRequestor(){ return new PythonSourceIndexerRequestor(); } public ISearchPatternProcessor createSearchPatternProcessor() { return new PythonSearchPatternProcessor(); } }[/size] Even I have PythonSourceElementRequestor and PythonSourceElementParser class implemented. Now, what else I need to do? I have doubt on indexing of my code. How can I print the index details? Does the indexing happens for all the file during the eclipse start up? ]]> birendra acharya 2012-09-17T12:21:40-00:00 Re: DLTK indexing <![CDATA[Hi, You'd better check JavaScript/Ruby/TCL implementations, as Python has some missing bits. Indexing happens on changes and also initial index update happens on startup. The data for the indexing are reported by *SourceElementParser - it's invoked with different requestors for building the module structure (Outline) and for the indexing. Search is performed in 2 steps: first source modules are identified based on index, then the matcher is called to find the exact matching nodes in the AST. Hope it helps. Regards, Alex]]> Alex Panchenko 2012-09-24T04:48:06-00:00 | http://www.eclipse.org/forums/feed.php?mode=m&th=377819&basic=1 | CC-MAIN-2016-40 | refinedweb | 257 | 50.33 |
I've managed to make a nice QML crossplatform app (Windows, Android) that uses a disconnected (runtime) local geodatabase. Now I need to add some Zip file extraction and the ability to test for a file's existence.
It looks like there are some perfect functions for that in ArcGIS.Extras ...
I can't find any reference that says how to install the QML Extras, and when I put this:
import ArcGIS.Extras 1.0
into my code, I get an error: QML Module not Found (ArcGIS.Extras).
Is there some trick to getting the Extras to work with AppStudio?
Thanks in advance for any help.
Please use the following import statement for AppStudio. All the components such as FileInfo, ZipFileInfo are under this library. No need to add "import ArcGIS.Extras 1.0"
import ArcGIS.AppFramework 1.0
Use the help within the QT creator for these components. There are some helpful code snippets on how to use these components with their properties and methods. Also the doc will specifically mention which import statement to use
Please post your AppStudio related issues here in future.
Thanks,
Nakul | https://community.esri.com/thread/196683-qml-arcgisextras-within-appstudio | CC-MAIN-2019-22 | refinedweb | 188 | 67.15 |
Handling of Different Injection Attacks in Grails
While implementing Security in my Sample Application I have read various types of Injection attacks that an application may suffer.
Reference: Grails In Action
1. SQL Injection Attack:
def username="gautam" Post.findAll(" from Post as post WHERE post.user.username='${username}' ")
This Query uses a local username property to control which posts are returned.
Try this Query in Grails Console.
An attacker can modifies the URL of the request so that the username parameter has the value :
def username = " ' or ' test' = ' test" Post.findAll(" from Post as post WHERE post.user.username='${username}' ")
The Query is Same, but this time username doesn’t look like an Id at all. Look what happens when we substitute the value into the query:
.. WHERE post.user.username = ' ' or ' test' = ' test'
Now all Post are returned, which will bring your server to a grinding halt.
By escaping input values before inserting it into the query, you can foil the attack.
The modified version of the HQL query that safe from the attack by escaping the value of username:
def username = " ' or ' test' = ' test" Post.findAll(" from Post as post WHERE post.user.username=? ", [username])
This is the Hibernate equivalent of a JDBC parameterized query.
2. Cross-Site Scripting (XSS) Attack
Another form of injection attack which targets HTML and javascript is when user Post this message
alert("alert")
A dialog pops up showing the message “alert”. Now every time you refresh your page, that message will pop up.
The solution of this is either:
– you can call the encodeASHTML() method on the text you want to display,
"${username.encodeASHTML()}"
But the implementation of the Grails tags like textField tag does the equivalent of encodeASHTML() method.
i.e,
<g:TextField
is equivalent of this:
attrs["value"].encodeASHTML()
An alternative is to use the defaultCodec page directive to enable HTML escaping on a page-by-page basis:
<% @ defaultCodec="html" %>
OR
-by adding/changing this entry in grails-app/conf/Config.groovy:
grails.views.default.codec="html"
By setting the default codec Grails uses to encode data in GSP views to HTML, you can ensure all GSP expression are HTML
escaped by default (This makes the Setting Global)
URL Escaping
Show Album
Simply by fiddling with the title parameter in a GET request an attacker could perform an XSS attack.To avoid this you can use encodeASURL() method on any data to be included in the URL.
Show Album
3. Other Form of Vulnerable Attack:
Alternative approach is to find out what platform the web application is based on. If you know that then you can narrow your hacking attempts to know the vulnerabilities of the platform.
There might be some other weakness in the Application like, try pointing your browser at this URL while your application is up and Running:
Any Attacker knows that the application is a java web application running on Jetty/Apache. Also,If application throws an exception ,Grails will display its standard error page. Then attacker also knows that your application uses Grails.
Solution to this Problem is mapping response codes to controllers like:
class UrlMappings{ "404" (controller:"errors",action:"notFound") "500" (controller:"errors",action:"internalError") }
But this mechanism can be bypassed, if you hard-coded to display the view like error.GSP if an exception is thrown by a GSP page, then it will declare to the user that your application is implemented with Grails.
You can modified your GSP to send a “500” error if environment is set to PRODUCTION:
${Response.sendError(500)}
Hope it Helps!!! | http://www.tothenew.com/blog/handling-of-different-injection-attacks-in-grails/ | CC-MAIN-2018-39 | refinedweb | 594 | 54.63 |
Created on 2016-03-18 10:23 by tds333, last changed 2016-04-08 22:10 by brett.cannon. This issue is now closed.
In site.py there is the internal function _init_pathinfo() This function builds a set of path entries from sys.path. This is used to avoid duplicate entries in sys.path.
But this function has a check if it is a directory with os.path.isdir(...). All this is fine as long as someone has a .zip file in sys.path or a zipfile subpath. Then the path entry is not part of the set. With this duplicate detection with none directories does not work.
The fix is as simple as removing the os.path.isdir(...) line and fixing the indent. Also the docstring should be modified.
Detected by using this function in a project reusing addsitedir(...) functionality to add another path with .pth processing.
Could you provide a code example of your using addsitedir that results in duplicates?
Here is a testcase to reproduce the issue:
> cat test.py
import site, sys
site.addsitedir('/foo/bar')
print (sys.path)
This prints just a single instance of '/foo/bar':
['']
Now if we explicitly set PYTHONPATH to include '/foo/bar'
> export PYTHONPATH=/foo/bar
and then run test.py here is the output:
['', '/foo/bar']
We see that there are duplicate entries for '/foo/bar'.
As Wolfgang rightly said the issue comes from the check for
"if os.path.isdir(dir)" inside _init_pathinfo() in site.py.
On removing this check I no longer see the duplicate entry for '/foo/bar'.
But since this is the first bug I am looking at I am not sure of the implications of removing this check. Can someone please confirm that what I see is indeed a failing test case, or is this the intended behavior?
Thanks,
Mandeep
Thanks.
Extended unit test for the issue and patch for site.py.
I think a fix for 3.6 only is ok, because it changes behaviour.
But this is only an internal function with a "_".
Should I add a test with a temporary created pth file?
Unfortunately you can't simply remove that directory check because you don't want to blindly normalize case. If someone put some token value on sys.path for their use that was case-sensitive even on a case-insensitive OS then the proposed change would break those semantics (e.g. if someone put a URL in sys.path for a REST-based importer).
The possibilities I see are:
1. Don't change anything; duplicate entries don't really hurt anything
2. Remove duplicate entries, but only normalize case for directories
3. Remove duplicate entries, but normalize case for anything that points to something on the filesystem (i.e. both directories and files)
And the code under discussion can be found at
I still think my fix is more appropriate as it ensures that known_paths and sys.path stay connected somehow.
Ok, I implemented point 3.
Check if it is a dir or file and makepath only in this case.
All other entries are added unmodified to the set.
Added a test case also for an URL path.
I think duplicate detection is now improved and it should break nothing.
New changeset bd1af1a97c2e by Brett Cannon in branch 'default':
Issue #26587: Allow .pth files to specify file paths as well as
New changeset 09dc97edf454 by Brett Cannon in branch '3.5':
Issue #26587: Remove an incorrect statement from the docs
New changeset 94d5c57ee835 by Brett Cannon in branch 'default':
Merge w/ 3.5 for issue #26587
I simplified Wolfgang's patch by simply using os.path.exists(). That eliminated the one place where os.path.isdir() was in use that was too specific to directories where files were reasonable to expect.
I also quickly tweaked the docs for the site module in 3.5 to not promise that files would work.
If you think there is still an issue with keeping things tied together, SilentGhost, feel free to open another issue to track it. | https://bugs.python.org/issue26587 | CC-MAIN-2017-13 | refinedweb | 676 | 77.33 |
How to use gazebo plugins found in gazebo_ros [ROS2 Foxy gazebo11]
I am trying to use the gazebo_ros_state plugin found inside the gazebo_ros folder in gazebo_ros_pkgs.
I have included this inside my world file similar to the demo also found in gazebo_ros folder...
<plugin name="gazebo_ros_state" filename="libgazebo_ros_state.so"> <ros> <namespace>/demo</namespace> <argument>model_states:=model_states_demo</argument> </ros> <update_rate>1.0</update_rate> </plugin>
I also tried to see if the migration guide found here (... and...) gave any info on it however nothing there seemed to offer any help.
This is the error I am getting when I try and start my world...
[Err] [Model.cc:1097] Model[cartpole] is attempting to load a plugin, but detected an incorrect plugin type. Plugin filename[libgazebo_ros_state.so] name[gazebo_ros_state]
libgazebo_ros_state.so is found inside my lib folder just like my other plugins
One more thing to note is that I have other plugins found in the gazebo_plugin folder that work and do not give an error like this...
I am having the same problem. Any updates here?
I took the gazebo_ros_state_demo.world straight from the repo. Strangely, I get the same error but regardless it seems to work and I get the two service endpoints /demo/get_entity_state and /demo/set_entity_state and they do correctly return model data.
I have one observation. The
gazebo_ros_stateplugin is registered as a WORLD_PLUGIN type here. The error in the OP is coming from
model.ccwhich is looking for
MODEL_PLUGINtype here.
So, it seems the problem here is, the addition of the plugin in the sdf file is not having the desired effect of specifying a WORLD_PLUGIN. I wonder if the
<plugin>tag added in the OP is under the
<model>or under
<world>. Not sure if that's a fix to the problem. But sharing a thought as I am trying to fix the same too. | https://answers.ros.org/question/356936/how-to-use-gazebo-plugins-found-in-gazebo_ros-ros2-foxy-gazebo11/ | CC-MAIN-2021-49 | refinedweb | 308 | 58.99 |
Two postfix operations in a single statement in GCC
#include <stdio.h> int z = 11; int main() { printf("%d\n", ((z++) * (z++))); printf("%d\n", z); return 0; }
$ gcc -o postfix_test.o postfix_test.c; ./postfix_test.o 121 12
Surprised? I sure was. It looks like gcc interprets two postfix operations in a single statement as a single postfix increment request. I guess this makes sense if you consider the postfix operator to mean, "Wait for this statement to complete, then have the variable increment." Assuming this specification, the second time that you postfix-increment the compiler says, "Yeah, I’m already going to have the variable increment when the statement completes — no need to tell me again."
On the other hand, prefix increment does work twice in the same statement. Maybe this is a decision that’s left up to the compiler? It’s not specified in K&R as far as I can see, but I haven’t checked any of the ANSI specifications.
Updates
2007/09/26 Here’s what Java has to say!
class DoublePostfixTester { public static void main(String[] args) { int z = 11; System.out.println(((z++) * (z++))); System.out.println(z); } }
$ javac DoublePostfixTester.java $ java DoublePostfixTester 132 13
Which is what I would have expected in the first place. Bravo, Java — we’re more alike than I thought. | http://blog.cdleary.com/2007/09/two-postfix-operations-in-a-single-statement-in-gcc/ | CC-MAIN-2017-30 | refinedweb | 223 | 67.45 |
= Fedora Weekly News Issue 97 = Welcome to Fedora Weekly News Issue 97 for the week of July 15th through July 21st 2007. The latest issue can always be found here[2] and RSS Feed can be found here[3]. To join or give us a feedback, please visit our project join page[4]. [1] [2] [3] [4] 1. Announcements 1. fedorapeople.org is now available 2. Smolt, Open Invitation 2. Planet Fedora 1. Ohio Linux Fest Keynote Address 2. GNOME Cookbook 3. Repoview-0.6.0 4. Mascots and Fedora. Do we need one? Do we want one? 3. Marketing 1. Volunteers needed for GITEX (8-12 September) 2. New in Fedora: Jack Aboutboul 3. Proposed Fedora 8 Features 4. Smolt to be a Linux Thing 4. Developments 1. Plans for tickless kernel for x86_64 architecture in Fedora 8 2. 'Allo 'Allo Wot's This 'Ere License? 3. Yum Integration For Applications 4. Java Based Web Interface To Fedora Repositories? 5. Sysklogd Replaced With Rsyslogd in Fedora 8 6. Presto-digitation 7. Seahorse: Reducing The Number Of Passphrase And Password Challenges 8. Nodoka Theme: Clean, Easy On The Eyes, Featured in Fedora 8 9. RUM RHUM RHUME REDRUM OPIUM OPYUM: Offline Fedora Package Manager 10. Another GNOME Conspiracy Unmasked: ShowOnlyIn 5. Infrastructure 1. Fedorapeople.org is up 6. Security Week 1. Computer Viruses are 25 Years old 2. Serious Security Issues in Samsung Linux Drivers 3. Firefox 2.0.0.5 Released 7. Daily Package 1. Krecipes - Recipe manager 2. Pulseaudio - Next-generation audio server 3. Crontab 4. Hwbrowser - Display hardware info 5. Ri-li - Run a wooden train 8. Advisories and Updates 1. Fedora 7 Security Advisories 2. Fedora Core 6 Security Advisories 9. Events and Meetings 1. Fedora Board Meeting Minutes 2007-07-17 2. Fedora Ambassadors Meeting 2007-07-19 3. Fedora Documentation Steering Committee 2007-07-17 4. Fedora Engineering Steering Committee Meeting 2007-07-19 5. Fedora Extra Packages for Enterprise Linux Meeting 2007-07-18 6. Fedora Infrastructure Meeting (Log) 2007-07-19 7. Fedora Packaging Committee Meeting 2007-07-17 8. Fedora Release Engineering Meeting 2007-07-16 9. Fedora Translation Project Meeting 2007-07-17 10. Extras Extras 1. LiveCD for Red Hat High == Announcements == In this section, we cover announcements from various projects. Contributing Writer: ThomasChung === fedorapeople.org is now available === SethVidal announces in fedora-announce-list[1], "What is fedorapeople.org[2]?:" [1] [2] === Smolt, Open Invitation === MikeMcGrath announces in fedora-announce-list[1], "Smolt[1] will reach 75,000 profiles in the next 24 hours and with that news I'm excited to announce functional clients that work in SuSE, Debian, and Ubuntu. With the help of the Linux community at large we could start to better understand what is out there. Look to changes in the near future like a ratings system, better reporting tools and other such improvements." [1] [2] == Planet Fedora == In this section, we cover a highlight of Planet Fedora - an aggregation of blogs from world wide Fedora contributors. Contributing Writers: ThomasChung === Ohio Linux Fest Keynote Address === MaxSpevack points out in his blog[1], "I've just purchased a plane ticket to Ohio Linux Fest[2], which is on Saturday September 29th. Joe Brockmeier is one of the primary organizers, and he asked me today if I would be willing to give one of the two keynote addresses. On behalf of all of Fedora, I am both flattered and excited to be given this opportunity." [1] [2] === GNOME Cookbook === JohnPalmieri points out in his blog[1], "A little birdy reminded me that GNOME's 10 year anniversary is coming up and I thought it would be nice to do something a little bit unusual to commemorate the creativity and passion which exemplifies our members. Since I have been talking about cooking classes and so many people in the GNOME community have also expressed their love for the culinary arts I thought it would be nice to publish a Creative Commons Attribution-ShareAlike 3.0[2] licensed cook book full of recipes from members of the GNOME community." [1] [2] === Repoview-0.6.0 === KonstantinRyabitsev points out in his blog[1], "Repoview[2] version 0.6.0 is available and should make life much easier -- this version introduces state tracking, so now repoview is aware of changes between runs. This means that it only generates and writes the files that have actually changed, which makes it extremely fast for small changes and also makes it rsync-friendly." [1] [2] === Mascots and Fedora. Do we need one? Do we want one? === NicuBuculei points out in his blog[1], "Back in the release cycle for Fedora 7 we had an initiative: create an open process where anyone can play and see if we can come with a Mascot for Fedora[2]. It generated a lot of long talks on mailing lists with strong supporters and opponents of the idea but also a number of contributions (about which I planned to blog but never came to it until now, shame on me!)." [1] [2] == Marketing == In this section, we cover Fedora Marketing Project. Contributing Writer: ThomasChung === Volunteers needed for GITEX (8-12 September) === JohnBabich reports in fedora-marketing-list[1], "I am doing my best to get a Fedora presence in the Red Hat EMEA booth at GITEX[2], being held this year on 8-12 September." "I can use all the help I can get to man the booth, answer questions, produce and distribute live CDs, etc. If you are attending GITEX this year and can spare some time to help out, I think it will be a productive and fun time for all involved." [1] [2] === New in Fedora: Jack Aboutboul === RahulSundaram reports in fedora-marketing-list[1], "Jack Aboutboul talks about what's new and interesting in Fedora in a Linux World podcast[2] with Don Marti. Why does the core/extras merge matter, custom spins of Fedora, relationship between Fedora and OLPC, KVM and virt-manager virtualization, does user space in Linux still suck like Dave Jones says it did, boot up speed and much more..." [1] [2] === Proposed Fedora 8 Features === RahulSundaram reports in fedora-marketing-list[1], "A pretty good look with screenshots and short descriptions though the list of features[2] are a bit premature to say at this point. I think we can do something similar to officially while announcing the list of features for Fedora 8 which will happen when we hit feature freeze during the development cycle." [1] [2] === Smolt to be a Linux Thing === MarcWiriadisastra reports in fedora-marketing-list[1], "Smolt, the Linux hardware profiler that was introduced by the Fedora project for automatically reporting installed hardware and other system attributes, reached a new milestone last week and is in the process of another. Last week, Smolt reached 75,000 profiles for Fedora after being introduced back in January of this year. At the time of writing, there are now over 78,300 profiles.[2]" [1] [2] == Developments == In this section, we cover the problems/solutions, people/personalities, and ups/downs of the endless discussions on Fedora Developments. Contributing Writer: OisinFeeley === Plans for tickless kernel for x86_64 architecture in Fedora 8 === (Editor's Note: This news beat was written by ThomasChung) WarrenTogami reports to FWN, "Hi Dave, Could you write up a paragraph describing Fedora's plan for x86_64 dyntick?" "We plan on releasing F8 with 2.6.23. Upstream seems against the idea of merging the 64bit tickless patches just yet, so will probably wait until 2.6.24 As a result of this, we'll carry patches in F8 to ship this feature early. Right now, as the 2.6.23 merge window is open, the tree is changing a lot, so adding patches at this stage would involve a lot of rediffing, so we'll remain without the tickless patches until things calm down when 2.6.23-rc1 is released. The plan of putting 64bit tickless in FC6/F7 updates has been put on hold until it's stabilised in rawhide, and may even wait until after F8 has been released, depending on how well testing goes. -- DaveJones" === 'Allo 'Allo Wot's This 'Ere License? === Following on from the Thursday, July 19th FESCo meeting[1] a request was posted[2] by BillNottingham to remind maintainers of the importance of keeping everyone informed about changing licenses. [1] [2] The request noted that even changes to license versions were important and that maintainers should ensure that the Fedora Project was notified about them via either of the lists: @fedora-devel-announce or @fedora-devel. Questions should be directed to FESCo. JakubJelinek wanted to know[3] whether the "License:" tags should be updated to reflect their exact versions now. JoshBoyer and BrianPepple answered that something such as the scheme which Jakub was proposing had been discussed in the FESCo IRC meeting[1]. That discussion was concerned with the technical aspects of how to provide license combinations in a compact, searchable space within current RPM strictures and TomCallaway (spot) expressed an objection to writing License tags of the form "License:GPL|MPL|BSD|X11|KitchenSink". There seemed to be plenty of practical objections to most of the suggestions and the meeting cohered around the proposal above, and also recognizing that much further work needed to be done. Dark murmurings about using a packagedb to hold the licenses were heard! [3] A slightly different answer was given by BillNottingham[4], who said that it was up to the Packaging Committee to standardize some naming convention and RalfCorsepius stated[4] that the Packaging Committee had already rejected versioned licensing due to considering the "License:" tag to be merely informative and not a legal statement, and to versioned licences introducing too much overhead. [4] Ralf followed up[5] on this with a strongly-worded riposte to Bill's original announcement, asking whether FESCo was now going to be the "Fedora license police". JefSpaleta thought[6] that Ralf's negativity was getting a bit too consistent and suggested the more positive construction which saw the development as a way of making sure that those that depend on particular packages aren't suddenly blindsided by a change to a license. Ralf gave further depth[7] to his objections, arguing that ignorance on the legal issues and a bureaucratic burden would hamper the Fedora Project's efforts to give substance to these proposals, he also dismissed GPLv3 as a consideration. JoshBoyer specifically countered[8] the latter point. ToshioKuratomi (abadger) agreed[9] that exotic licenses introduced complications, that the Packaging Committee had rejected guidelines for "License:" tags, but drew attention to the audience (end users) which had been considered in those deliberations. [5] [6] [7] [8] [9] === Yum Integration For Applications === Ignacio Vazquez-Abrams proposed [1] rewriting some system tools (e.g. ''authconfig'', ''system-config-network'', and ''desktop-effects'') to access ''system-install-packages'' so that they could install other packages "on the fly" to enable missing functionaility. [1] There was a muted reaction to the proposal. JochenSchmitt expressed[2] disquiet with the idea of doing something as intrusive as automatically installing a package on a running system without the explicit consent of an administrator. JesseKeating thought[3] that what was meant by "automatic" in this context was "automatically launch Pirut which would of course prompt for the root password." Further discussion between Jesse, Jochen, and Ignacio clarified that Ignacio was interested in using s-i-p with a Text User Interface instead of Pirut. [2] [3] JefSpaleta asked[4] whether a hand-created listing of packages for each tool would need to be made or if the process could be abstracted. Ignacio answered that s-i-p would be able to work exactly like yum in resolving needed packages and dependencies. [4] An alternate approach using "soft dependencies" (also discussed in the last paragraph of FWN#92 "Yelping Over Bloated Firefox And Flash"[4a]) was preferred[5] by KevinKofler. Kevin noted that this approach would avoid lock-in to yum/s-i-p. [4a] [5] MattMiller suggested[6] that a yum-plugin that allowed users to install what he termed "user level" applications using their own credentials would be useful. SethVidal was dubious, arguing that there was no easy distinction between user-level or other software. A detailed discussion[7] between HorstVonBrand and Matt over the dangers of Matt's suggestion versus the advantages of just using sudo followed and provided much food for thought. BennyAmorsen thought[8] that due to the difficulty of stopping users on UNIX systems making their own programs that the level of security which Horst wanted was already compromised. [6] [7] [8] === Java Based Web Interface To Fedora Repositories? === Recalling the use of repoview in the past, ThorstenLeemhuis wondered[1] what were the plans for a similar interface in the new merged Fedora and drew attention to a new project[2] named "Repowatch" run by RichardKörber (Shred). [1] [2] KonstantinRyabitsev (icon) responded[3] that he was completing work on a rewrite of repoview that had major speed-ups due to state-tracking. Icon pointed out that one advantage of repoview was that it did not require an engine to view the repository, being instead browsable simply as a collection of static pages. JesseKeating liked the sound of this and volunteered[4] to integrate into mash[4a] or pungi[4b]. [3] [4] [4a] Mash is a tool to query the koji buildsystem for particular RPMs and create a repository of them. [4b] Pungi is a set of python libraries that can be used to build composition tools. It is also a means of producing ISO images and/or installation trees. A response from the author/developer of repowatch clarified[5] that it was not a replacement for repoview, but instead provided a method to monitor data from sources such as repoview. RahulSundaram suggested requesting some Fedora Project resources officially (helpfully pointing to the place to do this) and ThorstenLeemhuis also encouraged[6] this, adding that repowatch could provide an easy way for users to find the latest versions of packages. [5] [6] JesseKeating noted[7] that it was possible to examine Koji to see what packages were available (see also FWN#88 "Making Koji A Complete rpmfind Replacement"[8]), but ThorstenLeemhuis was prepared for this and pointed back[9] to his earlier mail emphasising the need to cater to users. [7] [8] [9] Wishing to move things along practically, DavidTimms asked[10] whether Fedora Infrastructure could provide any parameters for Shred before he commenced a rewrite. Shred's response detailed his use of Tomcat leading to a negative reaction[11] from JesseKeating, who made it clear that Java apps (or PHP which was not under discussion here) were not something which he personally welcomed within the Fedora Project's essential infrastructure. A brief discussion of his reasons for this led Shred to decline[12] to enter the shopworn arguments about Java's performance and to conclude that there was no point in pursuing resource support within the Fedora Project. Jesse backed off from this conclusion and emphasized[13] that he had been merely expressing his own opinion. [10] [11] [12] [13] An interesting conclusion to the thread was provided[14] by NicolasMailhot who noted that the advent of Free Java and the importance of JBoss made it important for Fedora to be able to mount a credible working alternative to Microsoft's .NET stack and that Fedora Infrastructure should work with Red Hat/JBoss to mitigate any problems such as the ones Jesse claimed. [14] === Sysklogd Replaced With Rsyslogd in Fedora 8 === A replacement of the "sysklogd" kernel logging daemon was announced[1] by PeterVrabec. The reason[2] for this change is that sysklogd is no longer under active development. Very quickly there was disagreement over how this transition should be handled, with specific objections resting on the issues of what to call the the configuration files and whether the new daemon should be automatically started. [1] [2] MattMiller suggested that using rsyslog.conf instead of syslog.conf as the name of the configuration file replicated one of the problems which had been identified with syslog-ng. PeterVrabec thought that a simple ''cp syslog.conf rsyslog.conf'' took care of that problem, but ChuckAnderson pressed home[3] the point that a drop-in replacement ought to use the exact same names for configuration files. [3] A refinement of this point was made[4] by SethVidal who emphasized that it wasn't simple a replacement, but actually packaged as an "obsoletes". A direct response[5] from SteveGrubb (SteveG) stated that this had indeed been the original intent and that a sysconfig option had been originally present. SteveG also detailed some complicated hackery to conditionally use either of the filenames depending on the existence of sysconf.conf. There were some strong objections[6] from TomasMraz and KarelZak citing the simplicity of just settling on one name, using the ReleaseNotes to inform users of the change, and avoiding the use of a configuration file with a different base name to its daemon. LeszekMatok disagreed with the latter and posted[7] some examples during an exchange with RahulSundaram. MattMiller made a similar point elsewhere with a different example. [4] [5] [6] [7] The proposal to simply standardize on a new name was not congenial to DavidLutterkort who noted[7a] that it would require modification of all scripts which relied on the old names. SethVidal agreed and suggested[7b] a transition that would delay the final full switch until F9. [7a] [7b] RahulSundaram[8] and SethVidal[8a] voiced the other major objection: the case of users that upgrade using '''yum'''. Peter's announcement had made it clear that an '''anaconda''' upgrade to F8 would start the new rsyslogd daemon automatically, but otherwise an update would cause the old sysklogd to be erased and the new rsyslogd would need a ''su -c; /sbin/service rsyslogd start''. ManuelWolfshant thought[9] that leaving the system without a logger was enough of a problem to make an exception to packaging recommendations and start the daemon automatically from the %post section of the package. JeremyKatz objected[10] that this should only be done if sysklogd had been running previously or else there were negative effects for initing chroots, creating live-images, and installing some systems. [8] [8a] [9] [10] In response to ThorstenLeemhuis' suggestion to test whether sysklogd was running, JeremyKatz provided[11] an overview of how other scripts currently use a ''condrestart'' in %post but thought there would be problems depending on whether the sysklogd package had been removed before any tests could be run. BillNottingham amplified[12] this response, pointing out to MikeChambers that the PID file is what is examined. [11] [12] ManuelWolfshant concluded[13] that although there were potential problems for admins running other logging daemons for testing purposes, the proposed original scheme seemed mostly workable with the service starting after a reboot, and separately[14] JeffOllie noted that even in the case of "yum upgrade" from F7->F8 a reboot would be needed to use the new kernel and libraries anyway. [13] [14] === Presto-digitation === "Nodata" recalled[1] the discussion on including Presto (a way to reduce the amount of downloaded package data during updates by using diffs of RPMs) and wondered what was happening now, noting that the integration hadn't happened due to a lack of testing then and that the same thing seemed to happening again. [1] This summary was corrected[2] by DanHorák, who noted that contrary to Nodata's assertion there were Presto-enabled repositories in existence for FC6, F7, and Rawhide (the development branch). The generally favorable and pro-active approach of the Fedora Project to Presto had also been documented in FWN#93[3] and earlier. [2] [3] JeremyKatz added[4] that JonathanDieter (main Presto developer with AhmedKamal) had taken some patches from Jeremy and that the automatic generation of deltarpms into the buildsystem was being worked on. [4] Separately FESCo supported[5] the idea of fully integrating Presto into Fedora 8, with some minor worries being expressed as to whether fedora-infrastructure would be able to make the necessary changes in time for the Test1 freeze, and the clarification that as long as it made it by Test2 then it would make the cut. [5] A positive user experience with the existing Presto repositories was reported[6] by YuanYijun, with the caveat that a couple of errors needed to be worked around. FlorianLaRoche posted[7] an interesting link to an alternate GPL-licensed implementation of the binary-delta generating algorithm. [6] [7] === Seahorse: Reducing The Number Of Passphrase And Password Challenges === Following up on an earlier discussion, JesseKeating asked[1] whether it was possible to cut down the number of prompts for passphrases by managing ssh-agent passphrases within gnome-keyring. [1] RalfEtzinger explained[2] that OpenSSH currently uses a bundle of helpers (in openssh-askpass) and it ought to be easy to add a new one for gnome-keyring-ssh-askpass. [2] A suggestion[3] by ToddZullinger to use the pam_ssh module[4] provoked some minor debate as Todd admitted that this approach required the password entered to GDM (the display manager) to be the same as the SSH passphrase. JonathanUnderwood was opposed to the idea, but Todd argued[5] that enabling gnome-keyring to do what Jesse wanted would provide a similarly weak security model also compromisable at one single point. [3] [4] [5] AlexanderDalloz thought that "keychain" might be useful but wouldn't eliminate initial onetime password requests. This and the pam_ssh discussion led Jesse to clarify[6] that he envisioned a key storing application accessed by a single password which was different from each of the stored passwords/passphrases. GawainLynch pointed[7] out a Gentoo HOWTO and JonathanUnderwood advertised[8] SethVidal's Seahorse packages which apparently provide exactly the integration which Jesse had been seeking. [6] [7] [8] === Nodoka Theme: Clean, Easy On The Eyes, Featured in Fedora 8 === An announcement[1] of a new graphics theme for Fedora named "Nodoka" was posted by MartinSourada. Along with DanielGeiger, Martin aims[2] to provide a non-intrusive, consistent look throughout the desktop (exclusive of wallpaper which is release-dependent) using the "Echo" icons. [1] [2] Martin provided links to RPMs for those that wish to provide feedback and also sought further help in hosting the project. RahulSundaram followed up[3] on previous help he had given to Martin and suggested that the best course was to contact fedora-infrastructure to ask for resources. [3] It seems[4] that the theme will only be usable for GTK2 (as opposed to GTK1) applications (which hopefully are nearly eliminated at this stage) and KevinKofler wondered[5] whether someone was working to make this work usable with KDE as an alternative to Bluecurve/Quarticurve. Martin responded[6] that volunteers to do so were welcome. [4] [5] [6] The proposal that Nodoka be included in Fedora8 was later approved[7] by FESCo. [7] === RUM RHUM RHUME REDRUM OPIUM OPYUM: Offline Fedora Package Manager === (Editor Note: See for 'Naming Blues') One of the Google SoC projects was reported[1] to yield some usable results by Fedora users according to DebarshiRay (Rishi). Rishi has been working on a way to allow Fedora users without internet connections to benefit from the package management infrastructure already developed around yum. The result has been a tool[2] tentatively named "RUM", which exists to allow users to update a bandwidth poor machine by hooking it up to another machine that has updated packages on it. [1] [2] There was a certain amount of discussion about the name, with RahulSundaram drawing attention[3] to a conflict in namespace and querying some of the implementation choices, such as the use of uncompressed tar archives. Rishi seemed to settle[4] for "OPYUM" as the name. "Jima" commented[5] that as the contents of rpms were already compressed anyway there was probably not much gain in further compression. [3] [4] [5] A mild amount of skepticism about the value of this particular approach was expressed, especially by MikeChambers, who wondered[6] why Rishi couldn't just integrate his changes into Pirut. Rishi re-iterated[7] that the goal of the project was to allow even the extreme case of completely networkless machines being updated with a CD containing a "yumpack" specially built for them, and also explained that Pirut-maintainer JeremyKatz wasn't receptive to the idea so far. [6] [7] In response to Mike's wish for the ability of yum to update simply from a local network, ThomasSpringer provided[8] the exact command: '''su -c; yum update --disablerepo=\* --enablerepo=local-network''' [8] === Another GNOME Conspiracy Unmasked: ShowOnlyIn === An upset ChristopherStone asked[1] why so many applications were setting "ShowOnlyIn=GNOME" in their ''.desktop'' files and wondered if it was ''Fedora *trying* to cripple KDE or what⁈'' [1] RexDieter pointed out[2] the flaw in Christopher's approach and AlanCox suggested[3] that simple imitation of a flawed example rather than malice was a better working hypothesis and asked Christopher to file bugs if necessary. ToddZullinger posted[4] a revised grep, which indicated that GNOME users were actually victims of a KDE instigated desktop-cleansing campaign. [2] [3] [4] The plight of Xfce users was raised[5] by AndyShevchenko as they are affected by the preferential setting of the flag for both GNOME and KDE. [5] "TarekW" explained[6] that the advantage of the current setup is that it provides a sane default for non-power-users, shielding them from the potential bloat of having both desktops load their dependencies into memory. Power-users have the option to tweak ShowOnlyIn on a case-by-case basis. [6] == Infrastructure == In this section, we cover the Fedora Infrastructure Project. Contributing Writer: JasonMatthewTaylor === Fedorapeople.org is up === SethVidal and others have been working on the fedorapeople.org site for a while and their work is now available for use[1]. [1] == Security Week == In this section, we highlight the security stories from the week in Fedora. Contributing Writer: JoshBressers === Computer Viruses are 25 Years old === I ran across this story this week. The computing world is still suffering from computer viruses. They're a serious problem that waste a great deal of time and money. Given past trends, it's very likely that computer viruses will see their 50th birthday. === Serious Security Issues in Samsung Linux Drivers === This story is absolutely amazing It seems that by installing certain Samsung Linux drivers modify numerous applications on the local system, setting a number of them setuid-root. It's easy to claim open source produces higher quality software, this is a fine example of this. If the source is being scrutinized by the community at large, someone is likely to send a patch which can fix a plain stupid bug such as this. It's also quite likely that a developer will be more willing to "do it right" if the source is going to be available for the whole world to see. === Firefox 2.0.0.5 Released === A new version of the Mozilla products (Firefox, Seamonkey, Thunderbird) were released this week. It's rather important you ensure you've installed this update as the browser is certainly the most abused application these days. If you run Fedora, this update has been available via yum for several days. == Daily Package == In this section, we recap the packages that have been highlighted as a Fedora Daily Package. Contributing Writer: ChrisTyler === Krecipes - Recipe manager === ''Productive Mondays'' highlight a timesaving tool. This Monday[1] we covered Krecepies[2]: "Krecipes is a KDE recipe manager. It will store, seach for, and resize recipes, rate their nutritional content, and manage shopping lists. Recipes can be stored in plan files for personal access, or in MySQL or PostgreSQL databases for shared access (or very large recipe collections)" [1] [2] === Pulseaudio - Next-generation audio server === ''Artsy Tuesdays'' highlight a graphics, video, or sound application. This Tuesday[1] PulseAdudio[2] was featured: "PulseAudio is a next-generation audio server designed to be a 'Compiz for audio'. It enables you to start multiple audio applications (with different output systems) and direct each playback stream to the sink (destination) of your choice, even changing the destination on-the-fly, and to adjust the level of each individual channel." [1] [2] === Crontab === The ''Wednesday Why'' article[1] took a look at the ways that Fedora packages interact with the cron execution scheduler: "Packages that requires scheduled execution of jobs can be configued in either of two ways: they can include a crontab file which will be placed in /etc/cron.d (the approach used by the smolt package), or they can include a script file which will be placed in /etc/cron.hourly, /etc/cron.daily, /etc/cron.weekly, or /etc/cron.monthly (which is the approach used by cups)." [1] === Hwbrowser - Display hardware info === ''GUI Thursdays'' highlight a software that provides, enhances, or effectively uses a GUI interface. This Thursday[1], hwbrowser was discussed: "Hwbrowser is a very simple tool that provides read-only access to hardware information." [1] === Ri-li - Run a wooden train === ''Friday Fun'' highlights fun, interesting, and amusing programs. This Friday[1], we took a look Ri-li[2]: "Do you remember having a wooden train set when you were little? Ri-li is a neat little amusement that will remind you of those days." [1] [2] == Advisories and Updates == In this section, we cover Secuirity Advisories and Package Updates from fedora-package-announce. Contributing Writer: ThomasChung === Fedora 7 Security Advisories === * [SECURITY] blam-1.8.3-5.fc7 - * [SECURITY] bochs-2.3-5.fc7 - * [SECURITY] centericq-4.21.0-13.fc7 - * [SECURITY] chmsee-1.0.0-0.20.beta2.fc7 - * [SECURITY] clamav-0.90.3-1.fc7 - * [SECURITY] devhelp-0.13-9.fc7 - * [SECURITY] epiphany-2.18.3-2.fc7 - * [SECURITY] epiphany-extensions-2.18.3-2 - * [SECURITY] firefox-2.0.0.5-1.fc7 - * [SECURITY] galeon-2.0.3-10.fc7 - * [SECURITY] gimp-2.2.17-1.fc7 - * [SECURITY] kernel-2.6.22.1-27.fc7 - * [SECURITY] liferea-1.2.19-3.fc7 - * [SECURITY] nginx-0.5.28-1.fc7 - * [SECURITY] seamonkey-1.1.3-1.fc7 - * [SECURITY] thunderbird-2.0.0.5-1.fc7 - * [SECURITY] yelp-2.18.1-5.fc7 - === Fedora Core 6 Security Advisories === * [SECURITY] firefox-1.5.0.12-4.fc6 - * [SECURITY] gimp-2.2.17-1.fc6 - * [SECURITY] thunderbird-1.5.0.12-2.fc6 - == Events and Meetings == In this section, we cover event reports and meeting summaries from various projects. Contributing Writer: ThomasChung === Fedora Board Meeting Minutes 2007-07-17 === * === Fedora Ambassadors Meeting 2007-07-19 === * === Fedora Documentation Steering Committee 2007-07-17 === * === Fedora Engineering Steering Committee Meeting 2007-07-19 === * === Fedora Extra Packages for Enterprise Linux Meeting 2007-07-18 === * === Fedora Infrastructure Meeting (Log) 2007-07-19 === * === Fedora Packaging Committee Meeting 2007-07-17 === * No Meeting === Fedora Release Engineering Meeting 2007-07-16 === * No Meeting === Fedora Translation Project Meeting 2007-07-17 === * No Meeting == Extras Extras == In this section, we cover any noticeable extras news from various Linux Projects. Contributing Writer: ThomasChung === LiveCD for Red Hat High === RobinNorwood reports in fedora-livecd-list[1], "Last week, we held the second annual Red Hat High[2] here in Raleigh. I helped with the software side of things, and used the livecd tools to do it." "In a nutshell, it's a technology camp for incoming high school freshmen, using all open source software." "For the classrooms, we used lab space donated by NCSU. However, a couple of the labs were being used for NCSU classes during the week of the camp, so we couldn't just format the drives and install Fedora. Our solution was to use a live cd[3]." [1] [2] [3] -- Thomas Chung | https://www.redhat.com/archives/fedora-announce-list/2007-July/msg00010.html | CC-MAIN-2014-15 | refinedweb | 5,306 | 52.29 |
- Type:
Bug
- Status: Open
- Priority:
P2: Important
- Resolution: Unresolved
- Affects Version/s: 4.5.2, 4.6.0, 5.4.1, 5.12.0
- Fix Version/s: None
- Component/s: GUI: Printing
- Labels:
Previously logged and acknowledged as issue number #261362 in old system but never visible there or here.
I use QPrinter::setFullPage() to allow me to place a full-page sized template of a complex form (from an SVG file) on the page. On Linux it works fine. On Windows the template image is offset because, apparently, full page does not mean that on Windows (except when the print is directed to PDF).
This test code:
#include <QtGui> int main(int argc, char *argv[]) { QApplication app(argc, argv); QPrinter p(QPrinter::HighResolution); p.setPaperSize(QPrinter::A4); p.setOrientation(QPrinter::Landscape); p.setFullPage(true); qDebug() << "Page rect: " << p.pageRect(); qDebug() << "Paper rect: " << p.paperRect(); QWidget mainWin; mainWin.show(); return app.exec(); }
On Linux prints equal page and paper rectangles as the docs imply:
Page rect: QRect(0,0 14033x9917)
Paper rect: QRect(0,0 14033x9917)
and Windows with a 'real' printer selected:
Page rect: QRect(99,99 6814x4760)
Paper rect: QRect(0,0 7016x4961)
The result is the template is properly placed over the entire page in Linux but is offset right and down on Windows. If the default Windows printer is set to a virtual printer (e.g. Distiller or XPS Writer) the behaviour is correct.
Qt 4.5.2 on Linux. Qt 4.5.2 in Qt SDK 2009.03 on Windows
- is required for
QTBUG-25380 QtPrintSupport - Page Layout Issues
- Open
QTBUG-25384 QtPrintSupport - Windows issues
- Open | https://bugreports.qt.io/browse/QTBUG-5363 | CC-MAIN-2019-13 | refinedweb | 271 | 57.16 |
Up to Design Issues
This page assumes an imaginary namespace referred to as play: which is used only for the sake of example. The readers is assumed to be able to guess its specification.
Abstract: Natural languages, encodings, and similar relationships between one abstract thing and another, are best modeled in RDF as properties. I call these Interpretation properties in that they express the relationship between one value and that value interpreted (or processed in the imagination) in a specific way.
There has to date (2000/02) been a consistent muddle in the RDF community about how to represent the natural language of a string. In XML it is simple, because you never have to exactly explain what you mean. You can mark up span of text and declare it to be French.
His name was <html:span xml:Jean-Françla;ois</html:span> but we called him Dan.
Under pressure from the XML community to be standard, the RDF spec included this attribute as the official RDF way to record that a string was in a given language. This was a mistake, as the attribute was thrown into the syntax but not into the model which the spec was defining.
Consider the example in the identity section,
specifying that this thing is of type person, and has a common name, email address and home page as given.
Where to we add the language property? Of course we could add a language attribute to the XML, but that would be lost on translation into the RDF model: no triple would result.
Many specifications such as iCalendar (see my notes@link) would add another property to the definition of the person.
<rdf:description> <rdf:type></a> <play:name>Ora Yrjö Uolevi Lassila</play:name> <play:namelang>fi</play:namelang> <play:mailbox>ora.lassila@research.nokia.com</play:mailbox> <play:homePage></play:homepage> </rdf:description>
Here, the property play:namelang is defined to mean "A has a name which is in natural language B". In the iCalendar spec, the definition more complex in that the lang property is in same cases the language of a name and in other cases that of the object's description. This is a modeling muddle. The nice thing about doing it this way is that the structure is kept flat, and pre-XML systems such as RFC822 (email etc) headers have a syntax which can only cope with this.
There are many drawbacks to this muddle. Ora may have two names, one in Finish and another in English, and the model fails to be able to express that. Because the attribute is apparently tied to the person and not obviously attached to the name, automatic processing of such a thing is ruled out. Clearly, the structure does not reflect the facts of the case.
The second attempt is to make a graph which expresses the language as a property of the string itself. Clearly, "Ora Yrjö Uolevi Lassila" is Finnish, is it not? Yes, Ora is Finnish, but that is different. What we need to say is that the string is in the Finnish language. The problem, then, becomes that RDF does not allow literal text to be the subject of a statement. Never mind, RDF in fact invents the rdf:value property which allows us to specify that a node is really text, but say other things about it too. This is done by introducing an intermediate node.
<rdf:description> <rdf:type <play:name rdf: <rdf:value>Ora Yrjö Uolevi Lassila</rdf:value> <play:lang>fi</play:lang> </play:name> <play:mailbox <play:homePage </rdf:description>
There we have it, and in an RDF graph at least very pretty it looks. And indeed, we could work with this, apart from the fact that we have made another modeling error. It is not true that the language is a property of the text string. After all, the string "Tim" - is that English (short for Timothy? or French (short for "Timothé")? I don't need to add a long list of text strings which can be interpreted as one language or as another. A system which made the assertion that the string itself was fundamentally English would simply be not representing the case.
In fact, the situation is that Ora's name is a natural language object, which is the interpretation according to Finnish of the string "Ora Yrjö Uolevi Lassila". In other words, Finish the language is the relationship between Ora's name and the string. In RDF, we model a binary relationship with a property.
<rdf:description> <rdf:type></a> <play:name> <lang:fi>Ora Yrjö Uolevi Lassila</lang:fi> </play:name> <play:mailbox>ora.lassila@research.nokia.com</play:mailbox> <play:homePage></play:homepage> </rdf:description>
This works much better. Ora has a name which is the Finnish "Ora". This allows an RDF system to create a node for that string, and a "Finish" link from the concept of Ora the person, maybe a Danish link from the concept of the currency, and an old english link from the concept of weight (1/15 pound), not to mention a Latin link from the concept of the shore.
A problem we may feel is we would like the language to be a string, so that we can reference the ISO spec for all such things, but there is of course no reason why the spec for the lang: space should not reference the same spec.
Another problem we might feel is that it is reasonable for the play:name to expect a string, and in most cases it may get a string: what is the poor system supposed to do in order to accommodate finding a natural language object in place of a string? I guess making a class which includes all strings and all natural language objects is the best way to go. Any use of string which did not allow also such natural language object makes life much more difficult for multilingual software- so this is serious problem.
[[This leads us on to another interesting question of packaging in RDF. There is a requirement in XML packaging and in email packaging and it seems quite similarly in RDF that when you ask me for something of type X I must be able to give you something of type package which happens to include the X you asked for and also some information for your edification. But that is another story.@@@ eleborate and define properties or syntax@@@]]
What is really important is that we are using the ability of RDF to talk about abstract things, just as when we identified people by the resources they were associated with, but avoided pretending that any person had a definitive URI.
Datatypes here I mean in the sense of the atomic types in a programming language, or for example XML Datatypes (XML schema part 2). Defining datatypes involves defining constraints on an input string (for example specifying what a valid date is as a regular expression) and specifying the mathematical abstract individuals which instances of a type represent. One can model the relationship between the representation and the abstract value and the string using a property.
This doesn't tell us what it is 10 of. We could go through life without any model of types: we could define a shoe size as being a decimal string for a number inches. There are many questions and tradeoffs which datatype designers make (for example,
It would be nice to be able to model these questions in general in the semantic web, in order describe the properties of dat in arbitrary systems. We can introduce interpretation properties which link a string to its decimal interpretation as number, or a length including units. The problem is that the RDF graph which most folks use is the one above. The object of shoe:size is "10".
The simplistic system corresponding exactly to the Attempt 1 above, is to declare that shoe:size is of class integer. This implies (we then say) that any value is a decimal string. Given the string and the type we can conclude the abstract value, the integer ten. This works. It is the system used by XML datatytpes whose answers for the questions above are as I understand it [No, Yes, Yes, Yes, No]. A snag is that you can't compare two values unless you know the datatypes.
To model the representation explicitly in the RDF it seems you have to introduce another node and arc, which is a pain.
We can then define rdf:value to express that there is some datatype relation which relates the size of the shoe to "10". All datatype relations are subProperties of rdf:value with this system. Once it is that form, the datatype information can be added to the graph. You have the choice of asserting that the object is of a given class, and deducing that the datatype relation must be a certain one. You can nest interpretation properties - interpreting a string as a decimal and then as a length in feet. But this is not possible without that extra node. One wonders about radically changing the way all RDF is parsed into triples, so as to introduce the extra abstract node for every literal -- frightful. One wonders about declaring "10" to be a generic resource, an abstraction associated with the set of all things for which "10" is a representation under some datatype relation. This is frightful too you don't have "equals" any more in the sense you used to have it.
Instead of adding an extra arc in series with the original,
we can leave all Properties such as shoe:size as being rather
vague relations between the shoe and some string
representation, and then using a functional property (say
rdf:actual) to relate the shoe:size to a (more
useful) property whose object is a typed abstract value.
{ <#myshoe> shoe:size "10" } log:implies { <#myshoe> [is rdf:actual of shoe:size] [rdf:value "10"] } .
@@@ No clear way forward for describing datatypes in RDF/DAML (2001/1) @@
Interpretation properties was the name I have arbitrarily chosen for this sort of use. I am not sure whether it is a good word. But I want to encourage their use. Base 64 encoding is another example. It comes up everywhere, but XML Digital Signature is one place.
<rdf:description> <play:name <lang:fi <enc:base64>jksdfhher78f8e47fy87eysady87f7sea</enc:base64> </lang:fi> </play:name> </rdf:description>
Another example is type coercion. Suppose there is a need to take something of datetime and use it as a date:
<rdf:description> <play:event <play:start <play:date>2000-01-31 12:00ET</play:date> </play:start> <play:sumary>The Bryn Poeth Uchaf Folk festival</play:summary> </play:event> </rdf:description>
Such properties often have uniqueness and/or unambiguity properties. enc:base64 for example is clearly a reversible transformation. It it relates two strings, on printable and the other a byte string with no other constraints. The byte string could not in general be represented in an XML document. The definition of enc:base64 is that A when encoded in base 64 yields A. This allows any processor, given B to derive A. The specification of the encoding namespace (here refereed to by prefix enc:) could be that any conforming processor must be able to accept a base64 encoding of a string in any place that a string is acceptable.
Interpretation properties make it clear what is going on. For example,
<rdf:description <play:xml-cannonicalized <enc:hash-sha-1 <enc:base64>jd8734djr08347jyd4</enc:base64> </enc:hash-sha-1> </play:xml-cannonicalized> </rdf:description>
clearly makes a statement, using properties quite independently defined for the various processes, that the base64 encoding of the SHA-1 hash of the canonicalized form of the W3C home page is jd8734djr08347jyd4. Compare this withe the HTTP situation in which the headers cannot be nested, and the encodings and compression and other things applied to the body are mentioned as unordered annotations, and the spec has to provide a way of making the right conclusion about which happened in what order.
This pattern applies very well to units of measure.
See, for example a simple ontology of units of measure.
Representing the interpretation of one string as an abstract thing can be done easily with RDF properties. This helps make a clean accurate model. However, using the concept for datatypes in RDF is incompatible with RDF as we know it today.
See also:
@@@Needs circle-and-arrow pictures for each attempt.
Note. This section followed a discussion about "Using XML Schema Datatypes in RDF and DAML+OIL with DWC.
Thomas R. Gruber and Gregory R. Olsen, KSL "An Ontology for Engineering Mathematics" in Jon Doyle, Piero Torasso, & Erik Sandewall, Eds., Fourth International Conference on Principles of Knowledge Representation and Reasoning, Gustav Stresemann Institut, Bonn, Germany, Morgan Kaufmann, 1994. A non-RDF but thorough treatement including units of measure as scalar quantities.
Compare with SUMO units of Measure which seems have units as instances, and multupliers such as kilo, giga, etc as functions.
A ittle off-topic, On linear and area memasure, John Baez's "Why are there 63360 inches per mile?" is good reaing.
Up to Design Issues
Tim BL
(names of certain characters may have been misspelled to protect the innocent ;-) | http://www.w3.org/DesignIssues/InterpretationProperties.html | CC-MAIN-2016-07 | refinedweb | 2,227 | 60.55 |
Welcome to pstat, a freeware generalized framework for executing a lengthy operation in a thread.
Feedback is provided by a progress dialog which can optionally be cancelled. Included is a simple dialog based
application which shows how to use it. The program demonstrates the difference threading can make by showing
the difference with and without pstat. Try using the without pstat option from the test program and tab to
another program, when you tab back you will notice that it's window is white meaning that it is not responding
to any window messages. Now try the same with pstat. Notice the difference.
pstat
The sample application included with PStat allows you to calculate the 100,000th prime number both with and
without pstat. When executed without pstat you will notice that the sample application becomes totally hung
and fails to respond to window messages. With pstat you have no such problems.
When executed with pstat, while your function is being executed a progress dialog will be displayed as:
If you want to allow the user to cancel the operation during your function then the progress
dialog can be displayed with a cancel button as:
Features
History
v1.0 (27th March 1997)
v1.1 (18th February 1998)
CEvent
CProgressThreadDlg
v1.2 (8th November 1998)
Usage
To use pstat in your applications simply include pstat.cpp in your project and
#include "pstat.h" in whichever module wants to access it.
#include "pstat.h"
API Reference
PStat is made up of the single function as follows:
Return Value:
TRUE if the function was successfully executed otherwise FALSE i.e. the dialog was cancelled or some
other error occurred.
Parameters:
typedef ULONG (FUNCTION_WITH_PROGRESS)(void* pData, CProgressThreadDlg* pProgressDlg);
Remarks:
To see how your pfnFunction should be coded, have a look at the CalculatePrimeNumbers()
in pstattestdlg.cpp. As your function will be executed in the context of another MFC thread, you will
need to be aware of the usual synchronisation and access issues which come with multi-threaded MFC
applications.
CalculatePrimeNumbers()
Contacting the Author
PJ Naughter
Web:
8th Nove | http://www.codeproject.com/Articles/400/PStat-v1-2-Threading-Framework | CC-MAIN-2013-48 | refinedweb | 344 | 55.44 |
hey,
I'm trying to complete q3 from this pdf.
here is what ive written, so far it draws the middle line and the top box, but to me it looks like the label is off.
ive checked it and gone through it on paper and to me it should work fine (in theory), so now id like some help, see if someone else can spot my problem
here is my source. i have indicated the problem area in red.
import acm.graphics.*; import acm.program.*; import java.awt.*; public class ProgramHierarchy extends GraphicsProgram { private static final int BOX_WIDTH = 120; private static final int BOX_HEIGHT = 36; private static final int MIDDLE_LINE_HEIGHT = 36; private static final int GAP_BETWEEN_BOXES = 18; public void run() { placeMiddleLine(); placeTopBox(); } private void placeMiddleLine() { GLine middleLine = new GLine (getWidth()/2, (getHeight()/2) - (MIDDLE_LINE_HEIGHT)/2 , getWidth()/2, (getHeight()/2) + (MIDDLE_LINE_HEIGHT)/2); add(middleLine); } private void placeTopBox() { GRect topBox = new GRect ((getWidth()/2)-(BOX_WIDTH/2), ((getHeight()/2) - (BOX_HEIGHT)/2)-BOX_HEIGHT , BOX_WIDTH, BOX_HEIGHT); add(topBox); GLabel topLabel = new GLabel ("Program"); topLabel.setLocation((getWidth()/2) - (topLabel.getWidth()/2), (getHeight()/2) - (MIDDLE_LINE_HEIGHT/2) - (BOX_HEIGHT/2) + [COLOR="Red"](topLabel.getAscent()/2[/COLOR])); add(topLabel); } }
also, in the pdf it says this
and to me that last part doesnt make sense. to me it should read something likeand to me that last part doesnt make sense. to me it should read something likeThe labels should be centered in their boxes. You can find the width of a label by calling label.getWidth() and the height it extends above the baseline by calling label.getAscent(). If you want to center a label, you need to shift its origin by half of these distances in each direction.
im not sure if these two issues are directly connected, but help on either one would be appreciatedim not sure if these two issues are directly connected, but help on either one would be appreciatedIf you want to center a label, you need to shift its origin by half of these distances in each direction relative to... | http://www.javaprogrammingforums.com/whats-wrong-my-code/1460-centering-label-inside-rectangle.html | CC-MAIN-2015-11 | refinedweb | 337 | 56.29 |
HELLO BA!
first of all, thanks for your time and thanks for this webpage!, has been a endless resource of inspiration, help and learning material!
Im new here so here is a small introduction:
I’ve been around a month getting into game dev and, since i chose blender as a learning plataform, python.
what Im most interested into learning is coding, I started with C , but then kind of saw that python would engage me more as a beggining lenguage. Anyway, basics of modelling: done, basics of animating: done, basics of logic bricks: Done basics of python scripting: catastrofic (so far). hehe:RocknRoll:
anyway, the question in hand:
My point here is, for a empty, to move, then add a plane (1x1) on the spot, then move, then add a plane and so on, and then will end up with a 100X100 plain, so i scripted:
import bge
dunWidth = 100 #meters/tiles
dunHeight = 100 #meters/tiles
b = 0
a = 0
add= bge.logic.getCurrentScene().addObject
def main():
global b, a
cont = bge.logic.getCurrentController()
dunMarker = cont.owner
#Movement Calculation: (X, Y, Z)
while b < dunWidth:
b += 1
add(“FloorTile”, “DunMarker”,0)
dunMarker.applyMovement((1,0,0), False)
while a < dunHeight:
add(“FloorTile”, “DunMarker”,0)
a += 1
dunMarker.applyMovement((0,1,0), False)
#dunMarker.applyMovement((0,-dunHeight,0), False) #–> ment to erase the position of Y from 100 to 0 again, but dosent work.
main() but instead, to my surprise, it First add the tile, then goes through the loop ignoring the add(), so the result is a 1x1 tile at 0x0y and the empty ends at 100x100y… how many things Im doing wrong here?Aaaand, since we are here, how would you improve the coding?(trying to learn here
)
pd, yeah, Roguelike 3D project | https://blenderartists.org/t/python-misbehavior/648664 | CC-MAIN-2020-40 | refinedweb | 295 | 61.87 |
On 8/1/06, Jim Gallacher <jpg at jgassociates.ca> wrote:
> Deron Meranda wrote:
> > When you say req.user is writable, is that just for the member of the
> > Python "req" object, or does a write also modify the underlying
> > Apache C structure's request_rec->user (char*) member?
>
> request_rec->user is set.
>
> Here is the relavent bit from requestobject.c (3.2.10) in the
> setreq_recmbr function (line 1279):
> ...
> ...
>
> else if (strcmp(name, "user") == 0) {
> if (! PyString_Check(val)) {
> PyErr_SetString(PyExc_TypeError, "user must be a string");
> return -1;
> }
> self->request_rec->user =
> apr_pstrdup(self->request_rec->pool, PyString_AsString(val));
> return 0;
>
> Jim
Thanks Jim. That's what I was hoping it did.
I'm now down to thinking this must be a DAV or SVN thing. I can use
the builtin CGI handler along with my mod_python access handler and
can see that the request_rec->user is getting set (which shows up
as the REMOTE_USER environment variable in the CGI script).
# .htaccess
PythonAccessHandler setmyuser::accesshandler
SetHandler cgi-script
# setmyuser.py
from mod_python import apache
def accesshandler(req):
req.user = 'foobar'
return apache.OK
#!/bin/sh
# dump.cgi
echo "Content-Type: text/plain"
echo
echo "User is" $REMOTE_USER
exit 0
$ wget -O -
User is foobar
So somehow either mod_dav or mod_dav_svn must not be getting
this or need something more. More code diving....
--
Deron Meranda | https://modpython.org/pipermail/mod_python/2006-August/021705.html | CC-MAIN-2022-21 | refinedweb | 221 | 59.5 |
On 2005-07-22 15:57:03 +0100 Adrian Robert <address@hidden> wrote: > > On Jul 21, 2005, at 3:49 PM, Jeremy Bettis wrote: > >> I have a problem with the implementations of isEqual: and hash in NSDate. >> >> Let's say we have two dates: >> >> NSDate *a, *b; >> // This is actually a common case if you are doing floating point math to >> generate dates and get small rounding errors. >> a = [NSDate dateWithTimeIntervalSinceReferenceDate:100000.001]; >> b = [NSDate dateWithTimeIntervalSinceReferenceDate:99999.998]; >> >> printf("a = %d, b=%d, equal=%d\n", [a hash] , [b hash], [a isEqual:b]); >> >> // this code will print a = 100000, b = 99999, equal = 1 >> >> This breaks the NSDictionary rule that hash of two objects must equal if >> they are -isEqual:. >> >> I propose that we change the implementations to this: >> - (unsigned) hash >> { >> return (unsigned)([self timeIntervalSinceReferenceDate]+0.5); >> } >> - (BOOL) isEqual: (id)other >> { >> if (other == nil) >> return NO; >> if ([other isKindOfClass: abstractClass] >> && (int)(otherTime(self)+0.5) == (int)(otherTime(other)+0.5) ) >> return YES; >> return NO; >> } >> After my change the program's output changes to: >> a = 100000, b = 100000, equal = 1 >> >> I realize that the dates 100.5 and 100.49 are now not -isEqual:, but you >> have to draw the line somewhere, and putting it at .000 as the old code was >> worse. > > This really seems like a hack (not that it's any worse than the current state > :). Couldn't the implementations of -hash and -isEqual be aligned without > this loss of information? E.g. something like [warning, sloppy first-attempt > code here]: > > -isEqual:other > return abs(selfVal - other->selfVal) < epsilon; > > -hash > return selfVal / epsilon; > (or maybe (selfVal + epsilon/2.0) / epsilon) > > I think the problem with this is the division in -hash would need to be > carried out at a higher precision than the floating point representation of > selfVal uses. But maybe this could be worked around or lived with somehow? > E.g., if we used 8-byte floating point, epsilon was 0.001 (1 msec) and we > were willing to accept inaccuracy for dates beyond 100000 AD, would it work > then? Unfortunately my numerical computation class was WAY too long ago.. I agree that the current code is poor ... the implementation of -hash and -isEqual: must be compatibile with each other for a date to be used as a dictionary key, so you can't currently use dates as dictionary keys. However, I just checked with MacOS-X and the -isEqualToDate: method is explicitly documented as performing an exact comparison rather than just to the second. The original OpenStep specification on the other hand explicitly states that dates are equal if they are within a second of each other. Now, the OpenStep specification therefore rules out the possibility of using dates as dictionary keys. I therefore propose to follow MacOS-X, leaving the implementation of -hash as it is, and changing -isEqual: and -isEqualToDate: to perform exact comparisons. That allows us to use dates as dictionary keys and gives us MacOS-X compatibility. Comments? | http://lists.gnu.org/archive/html/gnustep-dev/2005-07/msg00099.html | CC-MAIN-2015-11 | refinedweb | 495 | 62.48 |
"Regan Heath" , dans le message (digitalmars.D:174462), a écrit :
> "Message-Digest Algorithm" is the proper term, "hash" is another, correct,
> more general term.
>
> "hash" has other meanings, "Message-Digest Algorithm" does not.
I think the question is: is std.hash going to contain only
message-digest algorithm, or could it also contain other hash functions?
I think there is enough room in a package to have both message-digest
algorithm and other kinds of hash functions.
On 8/8/12 8:54 AM, Regan Heath wrote:
> "Hash" has too many meanings, we should avoid it.
Yes please.
Andrei.
On Wed, 08 Aug 2012 14:50:22 +0100, Chris Cain <clcain@uncg.edu> wrote:
>.
I don't think there is any reason to separate them. People should know
which digest algorithm they want, they're not going to pick one at random
and assume it's "super secure!"(tm). And if they do, well tough, they
deserve what they get.
"std.digest" can encompass all message digest algorithms, whether secure
or not.
We could create a 2nd level below "secure" or "crypto" or similar if we
really want, but I don't see much point TBH.
R
--
Using Opera's revolutionary email client:
"Chris Cain" , dans le message (digitalmars.D:174466), a écrit :
>.
They should not be categorized the same. I don't expect a regular hash
function to pass the isDigest predicate. But they have many
similarities, which explains they are all called hash functions. There
is enough room in a package to put several related concepts!
Here, we have a package for 4 files, with a total number of line that is
about one third of the single std.algorithm file (which is probably too
big, I conceed). There aren't hundreds of message-digest functions to
add here.
If it where me, I would have the presently reviewed module std.hash.hash
be called std.hash.digest, and leave room here for regular hash
functions. In any case, I think regular hash HAVE to be in a std.hash
module or package, because people looking for a regular hash function
will look here first.
Am Wed, 08 Aug 2012 02:49:00 -0700
schrieb Walter Bright <newshound2@digitalmars.com>:
>
> It should accept an input range. But using an Output Range confuses
> me. A hash function is a reduce algorithm - it accepts a sequence of
> input values, and produces a single value. You should be able to
> write code like:
>
> ubyte[] data;
> ...
> auto crc = data.crc32();
auto crc = crc32Of(data);
auto crc = data.crc32Of(); //ufcs
This doesn't wok with every InputRange and this needs to be fixed.
That's a quite simple fix (max 10 lines of code, one new overload) and
not a inherent problem of the API (see below for more).
>
> For example, the hash example given is:
>
> foreach (buffer; file.byChunk(4096 * 1024))
> hash.put(buffer);
> auto result = hash.finish();
>
> Instead it should be something like:
>
> auto result = file.byChunk(4096 * 1025).joiner.hash();
But it also says this:
//As digests implement OutputRange, we could use std.algorithm.copy
//Let's do it manually for now
You can basically do this with a range interface in 1 line:
----
import std.algorithm : copy;
auto result = copy(file.byChunk(4096 * 1024), hash).finish();
----
or with ufcs:
----
auto result = file.byChunk(4096 * 1024).copy(hash).finish();
----
OK, you have to initialize hash and you have to call finish. With a new
overload for digest it's as simple as this:
----
auto result = file.byChunk(4096 * 1024).digest!CRC32();
auto result = file.byChunk(4096 * 1024).crc32Of(); //with alias
----
The digests are OutputRanges, you can write data to them. There's
absolutely no need to make them InputRanges as they only produce 1
value, and the hash sum is produced at once, so there's no way to
receive the result in a partial way. A digest is very similar to
Appender and it's .data property in this regard.
The put function could accept an InputRange, but I think there was a
thread recently which said this is evil for OutputRanges as the same
feature can be achieved with copy.
There's also no big benefit in doing it that way. If your InputRange is
really unbuffered you could avoid double buffering. But then you
transfer data byte by byte which will be horribly slow.
If your InputRange has an internal buffer copy should just copy from
that internal buffer to the 64 byte buffer used inside the digest
implementation.
This double buffering could only be avoided if the put function
accepted an InputRange and could supply a buffer for that InputRange so
the InputRange could write directly into the 64 byte buffer. But
there's nothing like that in phobos, so this is all speculation.
(Also the internal buffer is only used for the first 64 bytes (or less)
of the supplied data. The rest is processed without copying. It could
probably be optimized so that there's absolutely no copying as long as
the input buffer length is a multiple of 64)
>
> The magic is that any input range that produces bytes could be used,
> and that byte producing input range can be hooked up to the input of
> any reducing function.
See above. Every InputRange with byte element type does work. You just
have to use copy.
>
> The use of a member finish() is not what any other reduce algorithm
> has, and so the interface is not a general component interface.
It's a struct with state, not a simple reduce function so it needs that
finish member. It works like that way in every other language (and this
is not cause those languages don't have ranges; streams and iterators
(as in C#) work exactly the same in this case).
Let's take a real world example: You want to download a huge file with
std.net.curl and hash it on the fly. Completely reading into a buffer
is not possible (large file!). Now std.net.curl has a callback
interface (which is forced on us by libcurl). How would you map that
into an InputRange? (The byLine range in std.net.curl is eager,
byLineAsync needs an additional thread). A newbie trying to do that
will despair as it would work just fine in every other language, but
D forces that InputRange interface.
Implementing it as an OutputRange is much better. The described
scenario works fine and hashing an InputRange also works fine - just
use copy. OutputRange is much more universal for this usecase.
However, I do agree digest!Hash, md5Of, sha1Of should have an additional
overload which takes a InputRange. It would be implemented with copy
and be a nice convenience function.
>
> I know the documentation on ranges in Phobos is incomplete and
> confusing.
Especially for copy, as the documentation doesn't indicate the line I
posted could work in any way ;-)
Am Wed, 08 Aug 2012 11:27:49 +0200
schrieb Piotr Szturmaj <bncrbme@jadamspam.pl>:
> > BTW: How does it work in CTFE? Don't you have to do endianness
> > conversions at some time? According to Don that's not really
> > supported.
>
> std.bitmanip.swapEndian() works for me
Great! I always tried the *endianToNative and nativeTo*Endian functions.
So I didn't expect swapEndian to work.
>
> > Another problem with prevents CTFE for my proposal would be that the
> > internal state is currently implemented as an array of uints, but
> > the API uses ubyte[] as a return type. That sort of reinterpret
> > cast is not supposed to work in CTFE though. I wonder how you
> > avoided that issue?
>
> There is set of functions that abstract some operations to work with
> CTFE and at runtime:
>.
> Particularly memCopy().
I should definitely look at this later. Would be great if hashes worked
in CTFE.
> > And another problem is that void[][] (as used in the 'digest'
> > function) doesn't work in CTFE (and it isn't supposed to work). But
> > that's a problem specific to this API.
>
> Yes, that's why I use ubyte[].
But then you can't even hash a string in CTFE. I wanted to special case
strings, but for various reasons it didn't work out in the end.
>
> I don't think std.typecons.scoped is cumbersome:
>
> auto sha = scoped!SHA1(); // allocates on the stack
> auto digest = sha.digest("test");
Yes I'm not sure about this. But a class only based interface probably
hasn't high chances of being accepted into phobos. And I think the
struct interface+wrappers approach isn't bad.
>
> Why I think classes should be supported is the need of polymorphism.
And ABI compatibility and switching the backend (OpenSSL, native D,
windows crypto) at runtime. I know it's very useful, this is why we
have the OOP api. It's very easy to wrap the OOP api onto the struct
api. These are the implementations of MD5Digest, CRC32Digest and
SHA1Digest:
alias WrapperDigest!CRC32 CRC32Digest;
alias WrapperDigest!MD5 MD5Digest;
alias WrapperDigest!SHA1 SHA1Digest;
with the support code in std.hash.hash 1LOC is enough to implement the
OOP interface if a struct interface is available, so I don't think
maintaining two APIs is a problem.
A bigger problem is that the real implementation must be the struct
interface, so you can't use polymorphism there. I hope alias this is
enough.
Am Wed, 08 Aug 2012 11:27:49 +0200
schrieb Piotr Szturmaj <bncrbme@jadamspam.pl>:
>
> Yes, there should be bcrypt, scrypt and PBKDF2.
Wow, I didn't know about scrypt. Seems to be pretty cool.
On Wednesday, 8 August 2012 at 14:14:29 UTC, Regan Heath wrote:
> I don't think there is any reason to separate them. People
> should know which digest algorithm they want, they're not going
> to pick one at random and assume it's "super secure!"(tm). And
> if they do, well tough, they deserve what they get.
In this case, I'm not suggesting keep them separate to not
confuse those who don't know better. They're simply disparate in
actual use.
What do you use a traditional hash function for? Usually to turn
a large multibyte stream into some finite size so that you can
use a lookup table or maybe to decrease wasted time in
comparisons.
What do you use a cryptographic hash function for? Almost always
it's to verify the integrity of some data (usually files) or
protect the original form from prying eyes (passwords ... though,
there are better approaches for that now).
You'd _never_ use a cryptographic hash function in place of a
traditional hash function and vice versa because they designed
for completely different purposes. At a cursory glance, they bare
only one similarity and that's the fact that they turn a big
chunk of data into a smaller form that has a fixed size.
On Wednesday, 8 August 2012 at 14:16:40 UTC,
travert@phare.normalesup.org (Christophe Travert) wrote:
> function to pass the isDigest predicate. But they have many
> similarities, which explains they are all called hash
> functions. There
> is enough room in a package to put several related concepts!
Crytographic hash functions are also known as "one-way
compression functions." They also have similarities to file
compression algorithms. After all, both of them turn large files
into smaller data. However, the actual use of them is completely
different and you wouldn't use one in place of the other. I
wouldn't put the Burrows-Wheeler transform in the same package.
It's just my opinion of course, but I just feel it wouldn't be
right to intermingle normal hash functions and cryptographic hash
functions in the same package. If we had to make a compromise and
group them with something else, I'd really like to see
cryptographic hash functions put in the same place we'd put other
cryptography (such as AES) ... in a std.crypto package. But
std.digest is good if they can exist in their own package.
It also occurs to me that a lot of people are confounding
cryptographic hash functions and normal hash functions enough
that they think that a normal hash function has a "digest" ...
I'm 99% sure that's exclusive to the cryptographic hash functions
(at least, I've never heard of a normal hash function producing a
digest).
Am Wed, 8 Aug 2012 17:50:33 +0200
schrieb Johannes Pfau <nospam@example.com>:
> However, I do agree digest!Hash, md5Of, sha1Of should have an
> additional overload which takes a InputRange. It would be implemented
> with copy and be a nice convenience function.
I implemented the function, it's actually quite simple:
----
digestType!Hash digestRange(Hash, Range)(Range data) if(isDigest!Hash &&
isInputRange!Range && __traits(compiles,
digest!Hash(ElementType!(Range).init)))
{
Hash hash;
hash.start();
copy(data, hash);
return hash.finish();
}
----
but I don't know how make it an overload. See thread "overloading a
function taking a void[][]" in D.learn for details. | http://forum.dlang.org/thread/jvrjt6$qbj$1@digitalmars.com?page=4 | CC-MAIN-2014-42 | refinedweb | 2,163 | 67.35 |
I worked in multiple NodeJS projects that had to be connected to multiple databases and softwares at the same time. Every time I start a new project, I first need to write the code that configures clients of databases (MongoDB, ElasticSearch, Redis...), make sure it connected successfully and then move on to what I want to do.
The problem
The problem is that every client has it's own way to configure a client/connection, plus it's own way of checking if the connection was successful or not.
mongodbyou check with a callback(error, client) (supports Promises too).
elasticsearchyou initiate the client, then you going to need to call client.ping() to know if it works
redisyou need to listen to
connect,
errorevents
An idea
I need to make sure that I'm connected to all services before I start what I want to do. When I write code, I prefer working with Promises than callbacks, so I thought about wrapping the configuration step in a Promise that resolves the client/connection instance when it succeeds, and rejects the error when it fails like the example bellow :
import mongodb from 'mongodb' import elasticsearch from 'elasticsearch' import redis from 'redis' Promise.all([ // mongodb new Promise( (resolve, reject) => { mongodb.MongoClient.connect(mongodbURL, function (err, client) { if (err) reject(err) else resolve(client.db(dbName)) }) } ), // elasticsearch new Promise( (resolve, reject) => { var client = new elasticsearch.Client({ host: 'localhost:9200' }) client.ping({ // ping usually has a 3000ms timeout requestTimeout: 1000 }, function (error) { if (error) reject(error) else resolve(client) }) } ), // redis new Promise( (resolve, reject) => { var client = redis.createClient() client.on("error", function (error) { reject(error) }) client.on("connect", function (error) { resolve(client) }) } ) ]).then( ([mongodbClient, elasticsearchClient, redisClient]) => { // here I write what I want to do } )
The solution above worked for Me when I wrote scripts that are written in one file. I like to refactor my projects into multiple files/modules when it gets complicated, for example an API with
express that has multiple routes, I'de prefer to write them separately, it makes it easy to know where to look while debugging.
Now,
How am I going to access the clients from other files ?
With the
express example, we can use a middleware to include the clients in
req and access it in each route easily, but this is just an example of a project, what to do when we don't have middlewares as an option ?
To be honest you can figure it out, it depends on your project, what you want to do and how you going to do it, passing the clients as a parameter when you call other functions, pass them to constructors when you initiate objects, you're always going to need to decide where to pass them.
I'm a lazy developer, I want to focus on working on the solution, and I hate making it more complicated with the clients baggage. I wanted something that will be easy to setup and can be used everywhere !
Here is what I have decided to do:
the solution
I defined this interface to be followed while wrapping a database/software client
class DriverInterface { // methods // configureWithName is to support multiple configurations of the same software static configureWithName(name, ...clientOptions) // return Promise<client,error> // this just an alias that calls this.configureWithName('default', ...clientOptions) static configure(...clientOptions) // return Promise<client,error> // get client by name static getClient(name) // returns client // properties static get client() // an alias to this.getClient('default') static get clients() // returns all clients Map<string,client> }
I started with
mongodb and published it at npm as
@oudy/mongodb which can be used like this
Example
import MongoDB from '@oudy/mongodb' MongoDB.configure('test', 'mongodb://localhost:27017').then( database => { const users = database.collection('users').find() } )
Also if your project is refactored into multiple files/modules you can access the client using
MongoDB.client
Example
// models/getUsers.js import MongoDB from '@oudy/mongodb' export default getUsers(limit = 20, skip = 0) { return MongoDB.client .collection('users') .find() .limit(limit) .skip(skip) .toArray() }
Multiple databases
You can use
@oudy/mongodb to connect easily with multiple databases
Example
import MongoDB from '@oudy/mongodb' Promise.all([ MongoDB.configureWithName('us', 'myproject', 'mongodb://us_server:27017'), MongoDB.configureWithName('eu', 'myproject', 'mongodb://eu_server:27017') ]).then( ([US_region, EU_region]) => { // get from US US_region.collections('files').find().forEach( file => { // do our changes and insert to v2 EU_region.collections('files').insertOne(file) } ) } )
If you want access to the
us or
eu databases from other files you can use
MongoDB.getClient()
Example
// models/files.js import MongoDB from '@oudy/mongodb' export default getFiles(region, limit = 20, skip = 0) { return MongoDB.getClient(region) .collection('files') .find() .limit(limit) .skip(skip) .toArray() }
Now, what's next
I've implemented the same Interface on other packages
@oudy/elasticsearch,
@oudy/mysql,
@oudy/amqp,
@oudy/redis. I'm still working on documenting them properly.
We've been working with databases this way for 2 years now on multiple projects specially at CRAWLO (Leading a big data based software that helps e-commerce websites to increase their sales by monitoring internal and external factors).
I published the repository here github.com/OudyWorks/drivers. please check it out and consider contributing if you have suggestions or found bugs.
This is just one of the cool projects I've built (I think it's cool :D), based on this I've made other packages for building restful APIs, GraphQL servers and even web apps. They're already public in here github.com/OudyWorks (not documented yet). I'm planning to document them and write more articles to explain the story behind why I've made them.
I'm sorry for any typos you may find, this is my first time publishing an article about my work and I'm just so excited to share with you what I've been working on.
Please, feel free to leave comments bellow and follow Me if you're are interested in cools projects in NodeJS.
Discussion | https://dev.to/oudmane/multiple-databases-in-bigdata-projects-52n4 | CC-MAIN-2020-50 | refinedweb | 991 | 53.81 |
Write a C program to find the distance between two points. As per the Pythagoras theorem, the distance between two points, i.e., (x1, y1) and (x2, y2), is √(x2 – x1)2 + (y2 – y1)2.
This example accepts two coordinates and prints the distance between them. We used the C library Math functions sqrt and pow to calculate the square root and power.
#include <stdio.h> #include<math.h> int main() { int x1, x2, y1, y2, dtn; printf("Enter the First Point Coordinates = "); scanf("%d %d",&x1, &y1); printf("Enter the Second Point Coordinates = "); scanf("%d %d",&x2, &y2); int x = pow((x2- x1), 2); int y = pow((y2- y1), 2); dtn = sqrt(x + y); printf("\nThe Distance Between Two Points = %d\n", dtn); }
In this C program, the calcDis function accepts the coordinates of two points and returns the distance between those two points.
#include <stdio.h> #include<math.h> int calcDis(int x1, int x2, int y1, int y2) { return sqrt((pow((x2- x1), 2)) + (pow((y2- y1), 2))); } int main() { int x1, x2, y1, y2; printf("Enter the First Coordinates = "); scanf("%d %d",&x1, &y1); printf("Enter the Second Coordinates = "); scanf("%d %d",&x2, &y2); printf("\nThe Distance = %d\n", calcDis(x1, x2, y1, y2)); }
Enter the First Coordinates = 2 3 Enter the Second Coordinates = 9 11 The Distance = 10 | https://www.tutorialgateway.org/c-program-to-find-the-distance-between-two-points/ | CC-MAIN-2022-40 | refinedweb | 223 | 60.89 |
Reverse Engineering: The most powerful feature of Hibernate Tools is a database reverse engineering tool that can generate domain model classes and Hibernate mapping files, annotated EJB3 entity beans, HTML documentation or even an entire JBoss Seam application in seconds. With the help of Eclipse you can do reverse Engineering. Lets see step by step to reverse-engineer database tables to generate hibernate POJO classes and mapping XML files using hibernate-tools (eclipse).
About the example: I am using My Eclipse 8.6 and MySQL. We will generate POJO classes and mapping XML files etc and latter run a test program. Please see the self explanatory screen shots below.
1. Create a New Java project
File -> New -> Java Project
2. Add name of the project lets say HibernateExample and click Finish
3. Add Hibernate Capabilities to the project
right click on the project -> MyEclipse -> Add Hibernate Capabilities
4. Keep default settings as shown below and click Finish.
Note- Here we are using Hibernate 3.3
5. Default hibernate configuration file name is hibernate.cfg.xml. If you wish to change it else keep it as it is. Also do remember this is the file which will be referred from HibernateSessionFactory.java and used to create Hibernate Session.
6. Here add database credentials. Add driver, Database URL, Username and password. I am using MySql hence added JAR file mysql-connector-java-5.1.5.jar as shown below.
7. Add package name accordingly. I am creating new package 'com.javaxp.common' and class HibernateSessionFactory.java which will be used to create Hibernate Session. You can change it else keep it default as shown below and click Finish.
8. You can see below files are created.
HibernateSessionFactory.java
hibernate.cfg.xml
9. Now we will open Hibernate perspective.
Window -> Open Perspective -> MyEclipse Hibernate
10. You can add new database, if you are connecting for the first time else Just open the connection.
11. Provide the username and password.
12. Now you can see available tables under the schema.
13. Now creating packages 'com.javaxp.dao' to hold all DAO classes and 'com.javaxp.pojo' to hold all POJO classes.
14. Now we will do actual reverse engineering. Right click on the tables for which you want to create POJO classes.
right click on the tables -> Hibernate Reverse Engineering
15. Browse the POJO package, Click on Create POJO<>DB table mapping information and Java Data Object (POJO<>DB Table)
16. Here keep default settings
17. You can see all POJO classes and their respective hbm.xml files created.
18. Now create DAO classes as we did for POJO in step 14. Click on Java Data Access Object (DAO) (Hibernate 3 only), click next.
19. Now you can see DAO classes generated in package 'com.javaxp.dao'. Add import statements wherever required in DAO.
20. Now we willl create a test class i.e TestHibernate.java to test our project.
package com.javaxp.common;
import java.util.List;
import com.javaxp.dao.BsBooksDAO;
import com.javaxp.pojo.BsBooks;
public class TestHibernate {
public static void main(String[] args) {
BsBooksDAO bDao = new BsBooksDAO();
List
BsBooks book = null;
for(int i=0;i
book = allBooks.get(i);
System.out.println(book.getBookId()+" "+book.getBookTitle());
}
}
}
Run the program and you can see the output. | https://www.javaxp.com/2012/12/hibernate-reverse-engineering-demo.html | CC-MAIN-2019-30 | refinedweb | 546 | 53.37 |
Type: Posts; User: sclement
Dear all
I am slowly learning how to program wpf. One thing I am trying to do is plot a series of polylines on a canvas. I have gotton this far. However I would like the polylines to scale as I...
Dear all
I am slowly converting one of my apps to WPF, using model-view-presenter design pattern.
One thing I am trying to replicate is the plotting of multiple polylines of different colors...
Dear All
I am using several DataGridView's in a VS 2008 c# project. I set the defaultcellstyle in the properties pane, inluding the fore and background colors.
This seems to work ok...
Thanks for your help
cheers
simon
Thanks BigEd
I changed my code so that the arrays were no longer exposed. I'd be curious to see an example using IList
cheers
simon
Hi
I have a class which has several members which are arrays of doubles. I am trying to figure out the best way to access these arrays.
Is it best to use properties:
private static...
My boss at work, has asked me to make my 2 software products a bit easier on the eye, add some shading, fancy color schemes etc.
This is somehting I have always struggled with. Has anybody got...
Hi Matt
yes it was that pesky font size not being set to 100%. All working now
cheers
simon
Dear All
I've recently purchased a new laptop, with a different screen resolution (1920 x1080) to the one I do all my coding on at work (1680x1050). Ideally I would like to be able to transfer...
Hi Hannes
I thought that might be the case, as I could find no references on conversion when I did a search. Oh well just means a bit more coding !
cheers
simon
Dear All
I have adapted a C# application (written in VS 2005) to run along side a VB6 app. At the moment the VB 6.0 app dictates when the c# application is run.
Ideally what I would like to...
Hi Boudino & Cilu
thanks for the replies. I put some conditional statements in the assemblyinfo.cs. Like you state this changes the assembly title etc which is half of what I wanted to achieve. ...
Hi all
I have an GUI application written in c# VS 2005. I want to be able to to modify the assembly name and other assembly information depending on the context in which the application is built....
Hi Cilu
I tried that as follows:
[Serializable]
public class FilterParams : MarshalByRefObject
{
public Int32 param1 = 0;
public Int32 param2 = 0;
...
[VS 2005 , Vista]
Hi
I am having a bit of difficulty marshalling a structure which is the parameter in a unmanaged function call.
The structure is
[StructLayout(LayoutKind.Sequential,...
Hi Darwen
Thanks for the reply, it now works. Rather silly mistake on my part
thanks again
simon
Hi
I am working with a com component, which I have added a reference to in my C#. I am using windows vista (32 bit) and VS2005.
I have to implement a callback object which handles data being...
Got a solution to this
I have drawn my own progress bar, it actually looks ok when using a LinearGradientBrush, and it redraws when I want it to redraw
cheers
simon
hi
i am using a toolstripprogressbar in an application. I increment this using
PerformStep(), which is called on the tick event of a timer.
When I set the progressbar range from 0 to 9,...
Hi
I have quite a lengthy appplication that has a requirement to be localized into in Spanish as well as in English.
I have set the localizable flag to true on every form in the project, and...
Hi Torrud
it's now working. Unfortunately for no particular reason.
Strangely i have not seen this behavior under XP - perhaps is vista being flakey.
Hi Torrud
I should have added that I'm running under Vista.
I am pretty sure I deleted all copies of atomicdata.mdb from my hard-drive.
Interestingly I did a re-boot and and now my .exe...
This is fun !
I checked the connection string member of one of the table adaptors, displaying it in a message box.
It points to "c:\Program Files\company name\ product...
Hi
I have a c# application written with VS2005, that uses an access database.
When I run the application from within the the dev environment, it seems to connect to a different instance of the...
Hi
I have quite an old MFC project (which is now being built under VS2005), which has multiple language resources, for English, Polish, Spanish etc. The way I currently set-up which language... | http://forums.codeguru.com/search.php?s=ef1b51d038d739a83ef3f9d55c31ed5f&searchid=2757045 | CC-MAIN-2014-15 | refinedweb | 783 | 74.49 |
It's still the cover given by Xuemei
Directly enter the topic... What is used in vue is almost a process.
Introduction of Echarts
install
npm install echarts --save
Let's write an Echarts component and introduce it inside
import * as echarts from 'echarts'
Getting started
If you haven't contacted Vue or Echarts' partners before, the way to understand is undoubtedly Echarts official documents Or major search engines.
This is undoubtedly my way of understanding, but when I read the official documents, I think the example given by the official is a little inconsistent with Vue's style, but I'll post it first to realize the simplest introduction:
<template> <div id="echarts" style="width: 600px; height: 600px"></div> </template> <script> import * as echarts from 'echarts' export default { mounted() { this.createEcharts(); }, methods: { createEcharts() { // Initialize the ecarts instance based on the prepared dom var myChart = echarts.init(document.getElementById("echarts")); // Draw a chart // The most important thing is to understand the role of each configuration. If there is nothing to say, practice makes perfect myChart.setOption({ title: { text: "ECharts Getting started example", }, tooltip: {}, xAxis: { data: ["shirt", "cardigan", "Chiffon shirt", "trousers", "high-heeled shoes", "Socks"], }, yAxis: {}, series: [ { name: "sales volume", type: "bar", data: [5, 20, 36, 10, 10, 20], }, ], }); } } }; </script>
Just introduce this component into the app component. Let's see the initial effect.
problem
1) Change from DOM operation to ref
I wonder if you have found the problem:
In the official document, it is the document. Getelementbyid ("eckarts") that directly operates on the dom, which is actually different from Vue's concept.
In vue, we want to reduce direct dom operations as much as possible. How can we improve it here??
Let's clarify what document. Getelementbyid ("seals") gets. There is no doubt that the node information is obtained. When printed, you can see console.log (document. Getelementbyid ("seals");
Then we only need to obtain the node information in the way of vue, so we can use the ref attribute in vue.
<div id="echarts" ref="myEcharts" style="width: 600px; height: 600px"></div> console.log(this.$refs.myEcharts);
Let's take a look
It's as like as two peas we got before.
In order to be more consistent with the method in Vue, we will change this case to be more flexible:
<template> <div ref="myEcharts" style="width: 600px; height: 600px"></div> </template> <script> import * as echarts from 'echarts' export default { data() { return { myEcharts: null, option:{ title: { text: 'General chart' }, tooltip: {}, xAxis: { data: ["shirt", "cardigan", "Chiffon shirt", "trousers", "high-heeled shoes", "Socks"] }, yAxis: {}, series: [{ name: 'sales volume', type: 'bar', data: [5, 20, 36, 10, 10, 20] }] } } }, methods: { initChart() { this.myEcharts = echarts.init(this.$refs.myEcharts); this.myEcharts.setOption(this.option); } }, mounted () { this.initChart() } } </script>
That's how it's written in vue.
2) Optimization ideas
1. If we need to use multiple styles of charts in the project, it is not impossible to import them directly. However, if we only use the line chart or histogram, we still recommend that we import them on demand, so that the packaging volume of the project will be smaller.
On demand introduction of official documents
2. In addition, we can encapsulate echarts into components. This is the best way to make everything dynamic.
In my opinion, I only consider the optimized use mode, not the performance. If there are any deficiencies, please point them out in time. Thank you very much.
This is the simplest histogram. I know everyone's needs are different. Here's how to learn.
A variety of charts
In the official documents, there are many cases. Even if we don't know anything, the cv method can be easily used as soon as it comes out.
And each chart has core code.
The so-called core code is various configuration items in option. The official also gave a detailed introduction to each configuration item.
The column of data is data, and others are configuration items. However, how to render depends on our business.
Click the document column in the top menu, and there will be one Configuration item manual
The most commonly used configuration items should be the one in the figure below, commonly known as the nine configuration items.
If you want to understand it quickly, it is recommended to directly click on the official example diagram of a complex point, and then check it with the configuration manual. I think this method should be the simplest.
It is better to teach people to fish than to teach them to fish
Because different businesses may have different needs, we need to learn how to learn more.
You can draw so many kinds of pictures with Echarts. I can't make a small Demo for everyone, and everyone's data is different, so we can't copy our homework, so let's just talk about how to learn to play this thing.
Reading official documents in my opinion is always the fastest way to understand it. (maybe people will say why they don't read all kinds of blogs, or directly Baidu and Google)
There are two reasons:
- Because of the version, the blog you search does not necessarily use the latest version. The iteration of technology is really very fast now.
- Secondly, blog content is rarely as complete as official documents. In addition, in the process of writing, most authors (including myself) will post what they think is very core. For some very detailed things, they are likely to be ignored.
Hands on try
- After reading the document, it should not be applied to the current project immediately. The best way is to write a demo to see the effect (personal idea)
- I feel shallow on paper. I absolutely know that I have to practice it.
It's done. It's done | https://programmer.help/blogs/61a06c9ab8a8e.html | CC-MAIN-2021-49 | refinedweb | 966 | 63.7 |
So I have a saved game. when I launch the editor and open my pause menu screen and hit "load saved" everything works as expected. I am where i was and all the playerPrefs have loaded. Now if I play and die, "game over" appears and I get thrown back to the beginning of the level - as I want. Problem is that if I hit "load saved" from the pause menu now - without quitting out of play mode in the editor - I do go where I am suppose to and some of the saved game info works but some doesn't. Bottom line is i have to quit out of PLAY mode in the editor and then go into play mode and load the save level then . I guess it has something to do with the void Start() functions? that when entering PLAY mode they get activated making the "Load Saved" work then, but if I die and go back to the start and hit "Load Saved" it doesn't work because the Start functions aren't restarting? I don't know. DOes this make sense and how can I load a save point without stopping play mode?. I imagine in a build this would amount to having to quit out of the game app and restarting it to load a save point successfully. How do I reset all my scripts after death as well player position?
Answer by Lilius
·
Feb 23, 2018 at 07:37 PM
Maybe you could reload the scene:
using UnityEngine.SceneManagement;
public void RestartScene()
{
Scene thisScene = SceneManager.GetActiveScene();
SceneManager.LoadScene(thisScene.name);
}
Then load saved game data. You might want not to destroy some of your scripts and objects, use DontDestroyOnLoad:
Edit: Re-read the question and maybe this is what you are doing already? What happens when you die, do you reload the scene?
thanks....I actually stumbled onto something like this. I have a "restart" button in my pause menu as well - one i never even used until just a few minutes ago. and after I hit the restart button and was taken back to the start of the level and THEN hit "Load Saved" it worked without having to quit out of play mode and all the code said was
Application.LoadLevel(Application.loadedLevel);
Good. By the way, what version of Unity you are using? You should change that code to the one on my answer, because Application.LoadLevel is obsolete now and will not work if you update Unity.
I am using an ancient version 4.7. The game i am making is just a hobby that has been in the works off and on for a few years . I think I started in version 3. something. I have updated to 4.7 but that is as far as I go while working on this project. Tried not once but TWICE to upgrade to version 5 but it broke too many things - probably because of a lot of scripting code being obsolete - and simply not in the mood to back track and change code. Frankly a bit scared to upgrade even after I am done with this project as learning coding was a challenge and I'd hate to have to relearn the new changes. I have been making some changes just to eliminate some of the warnings. Like my graphics menu use to say QualitySettings.currentLevel = QualityLevel.Fantastic; and changed to the new and improved
QualitySettings.SetQualityLevel.
How do I reset my game but keep the Player in the same position?
1
Answer
R key= reset
3
Answers
Reset Timer after objects 've been destroyed
1
Answer
How Can I Reset Values?(Unity3D)
0
Answers
2
Answers | https://answers.unity.com/questions/1472574/reloading-a-scene-after-death-resetting-scripts.html | CC-MAIN-2019-35 | refinedweb | 617 | 80.11 |
Hi everyone again:P Here is some code for you newbies like me that want to use networking with your games, its simple and down the core of the program with a little chat pro-app. It just shows how a server can handle more then one client. You can open the client setup as much as you want.
Also for client put in your ipaddress (open cmd and type in ipconfig if you dont know it). If it sucks, then sorry lmao XD
Server:
import SocketServer class EchoRequestHandler(SocketServer.BaseRequestHandler): def setup(self): print self.client_address, 'connected!' def handle(self): while 1: data = self.request.recv(1024) print data if data.strip() == 'bye': return def finish(self): print self.client_address, 'disconnected!' #server host is a tuple ('host', port) server = SocketServer.ThreadingTCPServer(('', 5000), EchoRequestHandler) server.serve_forever()
Client:
import socket tcpstock = socket.socket(socket.AF_INET,socket.SOCK_STREAM) tcpstock.connect (("YourIpAddress", 5000)) while (1): data2 = raw_input('You>> ') tcpstock.send(data2) if data2 == "bye": tcpstock.close() print "You have disconnected!" break | https://discourse.panda3d.org/t/multi-client-handling/4445 | CC-MAIN-2022-27 | refinedweb | 168 | 61.83 |
Hello Every1,
I hope it is ok for me to post this thread here. I am new to VB .Net and am having difficulties knowing where to start with my new windows form application i am trying to build.
What i want to create is a form where a user can enter XML into a textbox and then hit 'TEST' to validate the XML to see if there is any problems with it.
I have built the form using VB express edition but now need to code and this is where i am coming across problems.
This is how the programe should work:
- A user selects a value from the combo box at the top of the form (This is the make of car. The xml structure changes for each value in the combobox)
- The user then enters the XML into the textbox or can do file open and select the XML which will then be displayed in this text box
- The user then hits 'Test' and the results of the XML are shown in another text box (Results Textbox) at the bottom of the form which highlight any problems with the XML or state if the XML is formatted correctly.
Here is a example piece of the XML that will need to be processed:
<File>
<Car>
<Vehicle Make="Ford">
<CarDetails Model="Fiesta"
Colour="Red"
Numberofdoors="5" />
<Owners Numberofowners="2"
Year
</Vehicle>
<Car>
</File>
I want the XML to be read so that all the attributes values entered are compared to a preformed list to check if it is a valid value entered.
For e.g If we look at the CarDetails node and the Model attribute, the value entered is Fiesta. I want my form to take this value and compare it against a list of valid values for this specific attribute, I.e Sierra, Escort, Fiesta. As Fiesta is on the valid entries list the program will continue reading the rest of the XML. If a value entered was 206 (A value that is not on accepted list) then the program continues to read the rest of the XML but highlights this problem in the results textbox. The user would then see what the problem is with the XML.
If anybody can help with this it would be very much appreciated as i am getting very frustrated!
Thank you!
Chloe ~X~
Welcome to the Forums!
This sounds suspiciously like a school project, but I'd recommend looking at the System.XML Namespace (and the XMLDocument class in it). In fact, the XML validation if extremely simple (as long as you don't mind only getting one error at a time...)
Code:
Dim A As New XmlDocument
Try
A.LoadXml(TextBox1.Text)
Catch Exception As XmlException
MsgBox(Exception.ToString)
End Try
As for other things, do some research into making schemas. The you should be able to define the correct values, and get the XML namespace to do all your work for you...
Dim A As New XmlDocument
Try
A.LoadXml(TextBox1.Text)
Catch Exception As XmlException
MsgBox(Exception.ToString)
End Try.
Hi Javajawa!
Thank you for this!
I have a quick question, how do i get the message to appear in a text box rather than a pop up message box? I have a second text box on my form called TextBox2 which i want the errors to be displayed in?
Also with this code it only looks for problems, so if the XML is formatted fine then there is no response, how can i add a message that says The XML is formatted correctly if there are problems with it?
Thanks Hun,
Chloe ~X~
You can set the text in a text box using TextBox2.Text =
As for displaying an ok message, the best Idea is to have a variable which counts the errors found. Then at the end of the test, you see if this is equal to zero. If it is, you display the 'It's All Ok' message...
Hi again,
I have set the TextBox2 as below but it doesn't display the error:
TextBox2.Text = (ex.Message)
Have i done something wrong?
How do i write a variable? Sorry for all the questions!
Chloe ~X~
Hi,
The TextBox2 message is working now!
I just want to write a variable to say that if there are no errors then display a message in TextBox2 stating The XML is formatted correctly, how do i do this?
Kind Regards,
Chloe ~X~
Don't worry about the questions - Everyone is a beginner at some point.
For the first part, I'm expect you're getting an error which isn't an XMLError, so it isn't processed in the catch section. The best way to check things is to set a breakpoint at the beginning of the procedure (F9) and step through and work out what is going on.
As for working with variables, Here is the MSDN page
I should pay more attention to replies, shouldn't I? :P
I'd recommend defining an Integer called 'ErrorCount' inside the procedure (have a read through the link in my last post if you're unsure). Then, inside the catch statement, add 1 to ErrorCount - this will then count the errors (yes, there may only ever be one, but something might change...)
Then, at the end of the procedure, see if it's 0, and change the text in the textbox accordingly.
Hi Javajawa!!
This is my code and it worked!
Dim xmlDoc As New Xml.XmlDocument
Dim errorCount As Integer = 0
Try
xmlDoc.LoadXml(TextBox1.Text)
Catch ex As Exception
errorCount = 1
TextBox2.Text = (ex.Message)
End Try
If errorCount = 0 Then
TextBox2.Text = "The XML is formatted correctly"
End If
Thank you for your help!!
What i need to do now is to get the program to check the values entered for each attribute in the XML is correct, so far i have this:
Dim xmlDoc1 As New Xml.XmlDocument
Dim xmlRoot As Xml.XmlNode
xmlDoc.LoadXml(TextBox1.Text)
xmlRoot = xmlDoc.DocumentElement
Dim make As String = xmlRoot.SelectSingleNode("Car/Vehicle").Attributes("Make").Value
Dim model As String = xmlRoot.SelectSingleNode("Car/Vehicle/CarDetails").Attributes("Model").Value
Dim colour As String = xmlRoot.SelectSingleNode("Car/Vehicle/CarDetails").Attributes("Colour").Value
Dim numberOfDoors As Integer = CInt(xmlRoot.SelectSingleNode("Car/Vehicle/CarDetails").Attributes("Numberofdoors").Value)
Dim numberOfOwners As Integer = CInt(xmlRoot.SelectSingleNode("Car/Vehicle/Owners").Attributes("Numberofowners").Value)
Dim yearRegistered As Integer = CInt(xmlRoot.SelectSingleNode("Car/Vehicle/Owners").Attributes("YearRegistered").Value)
How and where do i write the valid values that can be used in the XML?
For example the first atrribute value is 'Make' i want to be able to set 3 values that can be entered here, i.e AUDI FORD and VAUXHALL and so if BMW is entered this will return an error?
Thank you again so much!
Chloe~X~
May I ask where this is going? If this is some form of school work (and it does rather look like a school project), then all I can really do is to point you in the way of ENUMs, and Using the [ENUM].Parse method.
Also, you might want to store the car details as a structure.
View Tag Cloud
Forum Rules | http://forums.codeguru.com/showthread.php?466143-XMLReader | CC-MAIN-2017-39 | refinedweb | 1,204 | 72.26 |
Created on 2011-04-07.13:50:49 by pekka.klarck, last changed 2015-03-02.01:14:38 by zyasoft.
I got a report from a team using a tool I develop that the tool fails to start with this error:
IllegalArgumentException: Signal already used by VM: INT
This error occurs when the tool registers signal handlers with signal.signal. The code already catches ValueError which this method is documented to possibly raise [1] but obviously IllegalArgumentException goes through.
I looked at the source of signal module distributed with Jython 2.5.1 and noticed that it calls sun.misc.Signal.handle. I then looked at its documentation [2] which tells this in the beginning of the module doc:
"""Java code cannot register a handler for signals that are already used by the Java VM implementation. The Signal.handle function raises an IllegalArgumentException if such an attempt is made."""
It seems to me that calls to sun.misc.Signal.handle should be wrapped with try/except and possible IllegalArgumentException then reraised as a ValueError. I can provide a patch if this sounds like a good idea.
[1]
[2]
According to this StackOverflow question the underlying IllegalArgumentException problem is JVM specific:
Sounds good to me
A patch that wraps all sun.misc.Signal methods that may raise IllegalArgumentException with try/except that reraises them as ValueErrors is attached. As a result also signal.getsignal and signal.alarm may raise ValueError which they aren't documented to do, but that's probably better than raising equally undocumented Java based IAE.
I would have liked to create unit tests for these fixes but couldn't figure any easy way to do that. Probably tests could temporarily replace sun.misch.Signal.handle with a method that always raises an IAE but I'm not sure is that worth the effort.
IMHO, we should not be shipping or supporting anything in the sun.* namespace. Jython is sufficiently powerful that users are free and empowered to make use of these non-standard facilities if they wish.
But they should not be part of the core platform, which should rely only on java.*, javax.* or open source components in other namespaces. sun.misc has no place in jython.
For example, what happens if the code is run on a IBM JVM? Should we support sun.misc.* there?
If users wish to use JVM-specific facilities, they should be prepared to deal with the consequences.
My €0,02.
signal is the one case we use sun.misc services directlty in Jython, but they are also used in JFFI and Guava if available. However, this requires importing signal, so it's not an issue on unsupported platforms.
Fixed as of
Thanks for merging the patch Jim! If I would have implemented it now, I probably would have used a context manager like this instead:
from contextlib import contextmanager
@contextmanager
def map_java_exception(exception_map):
try:
yield
except tuple(exception_map) as err:
exp = exception_map[type(err)]
raise exp(err.getMessage())
with map_java_exception({IllegalArgumentException: ValueError}):
# ... | http://bugs.jython.org/issue1729 | CC-MAIN-2016-26 | refinedweb | 503 | 58.28 |
From: Geurt Vos (G.Vos_at_[hidden])
Date: 2001-04-09 09:38:56
First, sorry for not replying sooner. I just realized you were
working on this.
I've been working on a callback class for some time now, and
next to the obvious goals, there are two more requirements
I'm trying to meet:
1. the class should be extendable
2. it should be easy to extend
Both these criteria are not met by the any_function
implementation.
>
> ------------------------------
>).
Imagine the error message you'll get when the number of required
parameters isn't met, and any_function is called with either one
argument more or one less.......
> class any_function
why not 'class callback', and why only functions and functors,
and not also function members?
> any_function& operator=(const int); // assignment to
> zero clears target
This one seems quite superfluous. I mean,
just clear() seems enough to me.
> bool empty() const; // true if no target
> operator const void*() const; // evaluates true if non-empty
>
Again, I'm not certain whether both are required.
I personally would lose the operator...
Lastly, I'm not sure whether or not the distinction between
invoker and void_invoker is done for compatibility with
older compilers. Anyway, note that the following code is
legal C++ code:
void f1();
void f2()
{
return f1();
}
--------------------------------------------------
Up until now I only said what's 'wrong' with any_function,
of course I should also provide some solutions. First, I
can make my solution available, should I simply upload it
to groups.yahoo.com/boost?
Anyway, the declaration of the callback is as following
(omitted namespaces for brevity):
Callback<CxParam<Param0,...,Paramx,Result,validate> > instance;
'x' is the implementation for a specific number of
parameters. I have it implemented for 0 to 5.
Result defaults to 'void' and validate to 'true'.
If 'validate' is true, a call to operator() will
first check whether a function is attached and if
not, return Result();).
I'm not sure whether this all makes sense until you've
seen the code....just let me know where I should make it
available...
The two criteria I mentioned at the beginning of this mail
I think have been met quite nicely with this implementation.
I can make it even easier by using virtual inheritance at
two place, but I didn't want to touch that yet :)
Geurt
Boost list run by bdawes at acm.org, gregod at cs.rpi.edu, cpdaniel at pacbell.net, john at johnmaddock.co.uk | https://lists.boost.org/Archives/boost/2001/04/10848.php | CC-MAIN-2020-45 | refinedweb | 407 | 65.83 |
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(Cross
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A lot of customers have been asking about this so THANKS!!
You've been kicked (a good thing) - Trackback from SharePointKicks.com
How do you integrate Moss Menu into a MOSS 2007 publishing site?
How do you implement this control ? I have been looking for something like this but am a newb when it comes to javascript.
Hi Scott, the most natural place would be to edit your master page and replace an existing menu control (say, left nav) with your custom one. As with any control, the dll needs to be available on the server and needs to be registered as a safe control in the web.config.
How can we display the items in the menu based on some security rules.............say a particular menu item 'Site123' should not be visible to the users if some their user profile property does not meet a given criteria.
How can we have control on the menu items...i.e., I need to display /hide some menu items based on some user profile properties.
Does anyone know how to customize apperances of MossMenu?
My aim is I want to make MossMenu likes the navigation bars which is placed on.
The way I thinked is to render our custom style(HTML/CSS/JavaScript) or override appearence codes in MossMenu.
How to reuse those code to integrate into SharePoint 2007? By Web Part or User Control? I'm appreciated if anyone furnish any clue for this.
How can we have control on the menu items...i.e., I need to display /hide some menu items based on some user profile properties
- The MOSS 2007 navigation system takes care of trimming out items that users don't have permission to see, and you can also employ Audience Targeting to have some items visible to some users and not to others. If you want to do more advanced menu item trimming based on custom information this is probably most easily done by subclassing the MOSS navigation data source control (PortalSiteMapDataSource). For a good example of this, see the "Taxonomy and Tagging Starter Kit" post. It links to some code that that does exactly this in the TaxonomyPortalSiteMapDataSource class.().
- MossMenu is identical to the default ASP.NET 2.0 Menu control in terms of how it deals with CSS/styling. Look at the documentation for the System.Web.UI.WebControls.Menu control to learn about how to change the appearance using CSS.
We're having problems with the SharePoint Menu control not working after a post-back. Which javascript files does it rely on?
Dominion Digital , a Virginia-based Microsoft Gold Partner, recently launched the Performance Food Group’s
I have this problem with my WWS 3.0 and am not sure how to use this fix. Can anyone give me a step by step. instruction?
I am also having problems with this control. I have integrated it into the master page, but it does not seem to do one thing. The menu doesn't change, at all. I placed the script in the core.js file. Not sure what I could be doing wrong.
One fix that seemed to help was changing the asp:SiteMapDataSource to the SPSiteMapProvider and changing the attributes respectfully. Problem with this hack is, now it shows all sites on the topnavbar, regardless if I chose 'Display this site on the top link bar of the parent site?' when the site was first created. It is basically, showing the entire Site Map with Flyouts.
Does anyone have any suggestions or some simple directions on integrating this control into a WSS 3.0 site?
Thanks, in advance!
The product team has released the code for the menu control used in WSS v3 and MOSS 2007. This menu class
I would also like a step by step instruction on how to deploy this fix. I have created the dll and signed and deployed to the GAC, but SharePoint Designer does not recognize it.
Some basic steps would be helpful. -Thanks
Could any one point me to instructions to modify the menu code to allow images/buttons to bew displayed in the menu rather than text menu items?
Thanks.!
How to use the MossMenu
1. Download and install the "Microsoft Windows SharePoint Services 3.0 Tools: Visual Studio 2005 Extensions Version 1.0" on
2. Download MossMenu sources at
3. Create new microsoft visual studio 2005 project "Visual C#\Web Part"
4. Add MossMenu sources to the project
5. Add References to the project
System.Configuration
System.Design
System.Web
Windows Sharepoint Services Security (Version 12.0.0.0)
6. Change Default namespace ( Project > Properties > Application > Default namespace) in
Microsoft.SDK.SharePointServer.Samples
7. Sign the dll: Project > Properties > Signing > Sign the assembly
(key file: Properties\Temporary.snk)
8. Build using the "Release" Solution Configuration
9. Install the dll into the gac by copying the dll to the folder C:\WINDOWS\Assembly
10. Adding the control to the SafeControls list in the web.config file
Path: <configuration<SharePoint><SafeControls>
Add the following row:
<SafeControl Assembly="NL.ADA.MossMenu, Version=1.0.0.0, Culture=neutral, PublicKeyToken=9f4da00116c38ec5" Namespace="Microsoft.SDK.SharePointServer.Samples" TypeName="*" Safe="True" AllowRemoteDesigner="True"/>
Replace "NL.ADA.MossMenu" by the name of the dll.
11. Open the masterpage that uses the standard Sharepoint menu and register the new menu by adding the following tag at the top of the masterpage:
<%@ Register Assembly="NL.ADA.MossMenu, Version=1.0.0.0, Culture=neutral, PublicKeyToken=9f4da00116c38ec5" Namespace="Microsoft.SDK.SharePointServer.Samples" TagPrefix="Sharepoint" %>
12. In de masterpage replace the all <SharePoint:AspMenu> tags by the <Sharepoint:MossMenu> tag and the closing tags </SharePoint:AspMenu> by the <Sharepoint:MossMenu/> tag.
I have written a guide on how to improve the highlighting of the menu using this source.
Check it out:
In order to implements a custom menu based on this code, I've been trying to understand how to use the javascript code, but I'm not sure.... do I have to place the code in the script file into the core.js (in 12\TEMPLATE\LAYOUTS\1033) or there is anothe way ?
I'm newbie with MOSS stuffs but I thought it wasn't advisable to modify these files, and since our site must support localization, I understand that modifications to files into the 1033 directory (for En culture I think), will not be applied to other culture settings ?
Is that correct ?
I'd appreciate your help here !!
Well, with the guide up ist wrong.
The PublicKeyToken=9f4da00116c38ec5 cant work like this.
Anyhow, it looks pretty nice, but when it comes to run in the solution it doesnt work.
Building it np
Migrate that Big Issue.
Is there any other Informations about that?
send me a link @ timmey (at) live.de
i followed the steps but when i replace aspmenu with mossmenu in masterpage it gives error " object reference not set to instance of the object"
Is it possible to, Create Role based menu in moss 2007
I have tried the steps given by Roel.
On the sharepoint designer i am getting an error that says that "Object reference not set to an instance of an object".
And when i refresh the page the menu is also not rendered.
Could anyone help me on this. Please...
Thanks in advance..
SharePoint 2007 is a great platform for enabling collaboration among team members, and a huge step forward
When I try to implement this class and connect with the SPNavigationProvider on our site, it only displays a valid menu simetimes. When broken, some of the sub sites have no sub-nodes, just a menu item that says "Error" with an error stack in the Title property that reads:
An error occured while rendering navigation for requested URL: /dept. Exception message: Request failed. Stack trace: at Microsoft.SharePoint.Publishing.Internal.ServerResources.GetString(ResourceFileType fileType, String shortResourceName, CultureInfo culture)
at Microsoft.SharePoint.Publishing.PublishingWeb.get_PagesListName()Area.GetChildAreaIds()
at Microsoft.SharePoint.Publishing.Navigation.PortalWebSiteMapNode.PopulateNavigationChildren()
at Microsoft.SharePoint.Publishing.Navigation.PortalSiteMapNode.GetNavigationChildren(NodeTypes includedTypes, NodeTypes includedHiddenTypes, OrderingMethod ordering, AutomaticSortingMethod method, Boolean ascending, Int32 lcid)
at Microsoft.SharePoint.Publishing.Navigation.PortalSiteMapNode.GetNavigationChildren(NodeTypes includedHiddenTypes)
at Microsoft.SharePoint.Publishing.Navigation.PortalSiteMapProvider.GetChildNodes(PortalSiteMapNode node, NodeTypes includedHiddenTypes)
If I change the StartingNode property to "SID:1002" (which is what I really need it to do so it always displays the menu from the root), I don't get any menu at all unless the Master Page is changed--and even then, it's only partially there (the above menu is displayed.)
The regular SharePoint:AspMenu works just fine in our site...why won't this? I'd like to do some cool customizations with it, but not if it can't produce reliable menus.
If anyone has any ideas for me to try, please email me @ daniel.obrien(at)dmx(dot)com. I've been trying for about three weeks to build a control that will work on our site, and it just simply isn't working.
Many thanks guys...
-Dan
Is the javascript (Included in the zip file) need to be include on the page? or this is a part of core.js
Can I make any changes in design time HTML? I need to add some images in the tabs. How can i do that? | http://blogs.msdn.com/ecm/archive/2006/12/02/customizing-the-wss-3-0-moss-2007-menu-control-mossmenu-source-code-released.aspx | crawl-002 | refinedweb | 1,568 | 58.18 |
go to bug id or search bugs for
Description:
------------
To compile php_ini.c under QNX6.2 with GCC, I had to add the following lines (at the end of php.ini):
35: #include <sys/dir.h>
36: #include <sys/types.h>
Would it be possible to fix it for future release?
Thanks,
Alain.
Add a Patch
Add a Pull Request
In what file did you add those?
And what is HAVE_ALPHASORT defined in main/php_config.h ?
Sorry for the mistake, I add the files <sys/dir.h> and <sys/types.h> at the end of php.h (before #endif)!
in main/php_config.h:
#define HAVE_ALPHASORT 1
regards,
Alain.
What is HAVE_SYS_DIR_H defined in main/php_config.h ?
It's enought to include dir.h in php_scandir.h..
Just need to know it that define is there.
This line was commented out, I admit that I didn't see this definition BUT, I tried #define and #undef, that doesn't change anything!
The result of 'grep -in dir.h main/*.[hc] is:
main/internal_functions_win32.c:40:#include "ext/standard/php_dir.h"
main/php_config.h:661:/* Define if you have the <ndir.h> header file. */
main/php_config.h:662:/* #undef HAVE_NDIR_H */
main/php_config.h:715:/* Define if you have the <sys/dir.h> header file. */
main/php_config.h:716:#undef HAVE_SYS_DIR_H
main/php_config.h:733:/* Define if you have the <sys/ndir.h> header file. */
main/php_config.h:734:/* #undef HAVE_SYS_NDIR_H */
main/php_ini.c:33:#include "php_scandir.h"
main/php_scandir.c:28:#include "php_scandir.h"
main/php_scandir.c:41:#include "win32/readdir.h"
main/php_scandir.h:20:/* $Id: php_scandir.h,v 1.2.2.4 2003/02/19 18:45:03 sniper Exp $ */
main/php_scandir.h:22:#ifndef PHP_SCANDIR_H
main/php_scandir.h:23:#define PHP_SCANDIR_H
main/php_scandir.h:29:#include "win32/readdir.h"
main/php_scandir.h:50:#endif /* PHP_SCANDIR_H */
main/reentrancy.c:28:#include "win32/readdir.h"
As you can see, it seems that no file try to include sys/dir.h.
right?
Alain.
Get this:
And replacement files:
Replace the files, run configure and make.
If it fails, check again what HAVE_SYS_DIR_H is set to.
(this would be so much easier if I had access to QNX.. :)
Did the replacement files fix this problem?
Thanks,
I going to see that ASAP.
For QNX you can see, it should be possible to download the os for non commercial use, it's quite light.
Al.
everything is fine now.
thanks a lot,
Regards,
Alain | https://bugs.php.net/bug.php?id=25295&edit=3 | CC-MAIN-2020-29 | refinedweb | 410 | 55.91 |
!ATTLIST div activerev CDATA #IMPLIED>
<!ATTLIST div nodeid CDATA #IMPLIED>
<!ATTLIST a command CDATA #IMPLIED>
What is the suggested workflow is for working with 3DS=>Unity via FBX files?
If I update the FBX file and reimport in Unity, it's not the same as deleting the re-adding the FBX in Unity.
Deleting and re-adding works as one would expect, but this means any changes I've made to the prefab have to be redone (split animations, loop settings, removing erroneous geos, etc) - all links are lost etc; so this isn't a feasible workflow.
The simplest case right now is changing a texture. If I change a texture, overwrite the FBX file, and reimport it in Unity; it doesn't get updated. However, if I create a new prefab from that FBX file; the textures are the updated ones in the new prefab created by Unity.
So generally, what is the suggested workflow here? Why doesn't overwriting FBX files work as one would expect?
asked
May 24 '11 at 06:43 PM
v2k
14
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2
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2
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4
edited
May 25 '11 at 10:29 PM
Texture changes are not expected to work. It's by design. The idea is once you import model Unity creates materials and it doesn't overwrite them afterwards. It's done that way, because we want to allow users to edit their materials in Unity (setup right shader and so on) and we don't want to overwrite all that work when mesh is changed.
Meshes and animations should change as expected by just overwriting the file.
Regarding prefabs it's a known problem (i.e. if you add additional objects to the file they do not appear in the prefab). We have plans to improve our prefab system and these problems should be fixed then.
So to summarize: you should be overwriting your FBX files and it should work. If something doesn't - we want to know.
answered
May 25 '11 at 08:59 AM
Paulius Liekis
7.3k
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16
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24
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45
edited
May 25 '11 at 09:02 AM
But I want to change the materials I use in Max. How can I do this and expect them to update in Unity after I export to FBX and update that file in Unity? I am having this problem now and it is really frustrating.
It is my goal to touch FBX content files as little as possible in Unity. It is imperative that the workflow for me be smooth and seamless into Unity from Max.
The simple solution. Change a Max material, change its name and Unity is forced to pick up the change when you introduce the new FBX.
Also, simply over-writing FBX files already in the Assets dir seems to sometimes be problematic. Particularly if you already have them under source control. I export them to a separate temp dir and then copy them into Assets. Seems to produce more consistent and desirable results.
answered
Aug 06 '12 at 02:54 PM
diggerjohn
2
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2
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2
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6
edited
Aug 06 '12 at 01:47 PM
I am having the same trouble with materials. I want to update the materials through max. So how can I update the FBX and expect the materials to change as well as geometry?
answered
Aug 03 '12 at 09:22 PM
edited
Aug 03 '12 at 07:33 PM
Do not post questions as answers. There are comments for that.
I am having the same problem with materials. How can I make changes in materials in max and have it reflected in Unity through FBX updates?
answered
Aug 03 '12 at 07:44 PM
edited
Aug 03 '12 at 07:34 PM
Binaries and ASCII are fragile in complicated projects or when we make it complicated. I would plan the project properly before importing assets into unity which would establish smooth progress and pleasant workflow rather than having a go in hard ways.
My solution was to simply change the name of the Material in Max. This forced Unity to recognize it when I copied in the new FBX into the Assets dir.:
import x969
fbx x556
3dsmax x334
asked: May 24 '11 at 06:43 PM
Seen: 1754 times
Last Updated: Aug 06 '12 at 02:54 PM
performance (RAM usage&FPS drop) difference between FBX and max files
Animation from 3dsMax to Unity
3DS Max & Unity - The best way to import a Multi/Sub-Object Material?
3ds max could not be found (.max asset in project)
"Max couldn't convert the max file to an FBX file!" error
FBX Importer Problems
ASE Import or writing an ASE Importer
How to make a player by 3dmax?
Exporting from 3DSMax 8...
Geometry and color are not conserved between 3DS Max fbx conversion and Unity
EnterpriseSocial Q&A | http://answers.unity3d.com/questions/121209/fbx-workflow-3dsgtunity.html | CC-MAIN-2013-20 | refinedweb | 806 | 72.36 |
Is there a script ?
What about the sendmail perl module ?
Any ideas ???? :)
Thanks in advance
Dazz
You don't need a local server to send mail from your box.
I am doing another hobby site :) and I want to send emails without using the Yahoo SMTP because of the footer.
I just want to be able to send emails without any footer, or a footer I can control.
Thanks for your help
If you're hosted on a Windows box, things will look different. I'm not familiar with any standard mail injection mechanisms there, if those even exist. If you don't find anything, using a SMTP server might indeed be an option. But then, that doesn't necessarily have to be one you run yourself, not even one run by your hosting provider.
In other words, your problem will be easier to solve when you tell us a little more about the infrastructure you have at your disposal... ;)
Apache, Perl 5.6, sendmail and whatever else I need ;)
Thanks
this is the message text that may or may not go over several lines
Then invoke "/usr/sbin/sendmail -t", probably with whatever implementation of popen you have available in Perl, and feed it the above message on stdin. The first empty line in the data stream seperates the headers from the message text (you could add other headers like CC: or BCC: if you need to). The "-t" option to sendmail tells it to take the recipients and sender addresses from the headers in the data. In pseudocode (or Python), a working procedure would look like this:
def send_with_sendmail(recipient, sender, msg): f = os.popen("/usr/sbin/sendmail -t", "w") f.write("To: " + recipient + "\n") f.write("From: " + sender + "\n") f.write("Return-Path: " + sender + "\n") f.write("\n") f.write(msg) f.write("\n") f.close()
I will be busy for a while I think :)
I'll give it a try, thasnks for the pointers. | http://www.webmasterworld.com/forum23/793.htm | CC-MAIN-2014-52 | refinedweb | 328 | 73.17 |
Cutting a long story short, it normally should. Unless you are using
ConfigureAwait(false) which can have a side effect with continuation not
flowing the context.
Alternatively try adding this setting in your app.
<appSettings>
<add key="aspnet:UseTaskFriendlySynchronizationContext" value="true"
/>
</appSettings>
UPDATE
NOTE!!
Initially I put false. But it must be true so that context flows.
HttpContext Encapsulates all HTTP-specific information about an
individual HTTP
request.
Hence each request HttpContext.Items["sameKey"] will be a different copy.
An ASP.NET Page is also an HTTP Handler through IHttpHandler interface
implementation and therefore is a valid candidate to use with the
HttpContext.Items collection by definition in the MSDN document.
Because Page Controls are effectively in a handler there seems no apparent
risk to reference page controls from the Items collection.
My main concern was lifetime of control objects in context of the Page
versus longer lifetime of the request/context the collection relates to.
However Controls will still be disposed properly through IDisposable when
the page is complete and the garbage collector will still clean up Controls
when they are finally released from the Items collection which shouldn't be
long after the page handler expires.
I figured it out. I was also using a custom compression attribute I had
forgotten about (thanks, @Steve) which I presume was clashing with the
Ombres settings, or at the very least, not decompressing properly. I
removed those attributes anyway and the normal error pages turned up, which
coincidentally related to Combres.
I stay away from this world by:
Changing the folder structure to this:
.MySolution.sln
.MyFirstProjectMyFirstProject.csproj
.MySecondProjectMySecondProject.csproj
.MyThirdProjectMyThirdProject.csproj
.MyOtherProjectSomeFolderMyOtherProject.csproj
.ThirdPartyReferences
Put all of your referenced assemblies in ".ThirdPartyReferences".
Have all the csproj reference the .dll's using the
"..ThirdPartyReferencesMyCoolDll.dll"
(however many ".." you need)
The every csproj has to use the same version and originating location of
any third party references.
Yes you are in callback hell. The solution assuming you don't want to use
async (which I doubt you can justify other than prejudice) consists of:
1) Make more top-level functions. Each function should perform either 1 or
2 IO operations as a rule of thumb.
2) Call those functions, making your code follow a pattern of a long list
of short core functions organized into business logic by a small list of
control flow "glue" functions.
Instead of:
saveDb1 //lots of code
saveDb2 //lots of code
sendEmail //lots of code
Aim for:
function saveDb1(arg1, arg2, callback) {//top-level code}
function saveDb2(arg1, arg2, callback) {//top-level code}
function sendEmail(arg1, arg2, callback) {//top-level code}
function businessLogic(){//uses the above to get the work done}
3) Use more functi
According to Wikipedia:
The byte order mark (BOM) is a Unicode character used to signal the
endianness (byte order) [...]
The Unicode Standard permits the BOM in UTF-8, but does not require nor
recommend its use.
Anyway in the Windows world UTF8 is used with BOM. For example the standard
Notepad editor uses the BOM when saving as UTF-8.
Many applications born in a Linux world (including LaTex, e.g. when using
the inputenc package with the utf8) show problems in reading BOM-UTF-8
files.
Notepad++ is a typical option to convert from encoding types, Linux/DOS/Mac
line endings and removing BOM.
As we know that the UTF-8 non-recommended representation of the BOM is the
byte sequence
0xEF,0xBB,0xBF
at the start of the text stream, why not removing it with R itself?
## Conv
One option might be to delete:
$this->connection->autocommit(FALSE);
$this->connection->commit();
in every function that is relevant (a,b)
and then wrap them around your function calls.
class MyClass {
private $connection = ...; // mysqli connection
public function a () {
// do something and run multiple queries
}
public function b () {
// do something and run multiple queries
$this->a(); // running function a inside function b here
}
}
// to run the code:
$myclass = new MyClass();
$myclass->connection->autocommit(FALSE);
$myclass->a(); // all queries are run inside a transaction
$myclass->b();
$this->connection->commit();
UPDATE:
OR maybe better , use a wrapper-function for calling the functions inside t
It looks like an encoding problem. Your application uses one encoding,
while the server uses other.
Using the Charset class will be your answer. Use it when converting
received data to String. Most probably you'll have to specify it in a
Reader constructor, though I can't say without any code.
Here is the link to appropriate javadoc: InputStreamReader(InputStream,
Charset)
You have a WHERE clause in your INSERT statement which makes no sense. If
you are inserting, you create a new row. I suspect you intend to UPDATE an
existing row in which case your SQL should be:
"UPDATE PropertyInfo SET FePool = @0, FeJac= @1, FeGames = @2 WHERE
PropertyID=@3"
If a checkbox is not checked, nothing will be included for that checkbox in
the Request.Form collection. If it is checked, the value passed is "on" by
default, or whatever you have specified in the "value" attribute. For
check1, that will be "true". For check2, that will be "check2". You need to
establish whether the checkbox was actually included in the Request.Form
collection and then set the values you want to commit to the database
accordingly. If you want to pass true or false depending on whether the
checkb
The error is in SystemServiceManager.java of Holoeverywhere. If you read
the code you'll noticed that it is intentionally throwing the error as
something is going wrong with annotations
public static void register(Class<? extends
SystemServiceCreator<?>> clazz) {
if (!clazz.isAnnotationPresent(SystemService.class)) {
throw new RuntimeException(
"SystemServiceCreator must be implement SystemService");
}
SystemService systemService = clazz.getAnnotation(SystemService.class);
final String name = systemService.value();
if (name == null || name.length() == 0) {
throw new RuntimeException("SystemService has incorrect name");
}
MAP.put(name, clazz);
}
A check of the proguard troubleshooting document shows that by default i
In case all your #define's are in a single configuration file, you can try
doing a preprocessor-only compilation and then finding all defined macros.
E.g.,
$ gcc -DFEATURE1 -DFEATURE2 -E configuration.h | grep '#define'
Why those packages installed in those locations?
/usr/local/lib/ghc-7.6.3 contains libraries that come with ghc. And
/usr/local/lib contains libraries from "haskell platform"
How do i delete the whole cabal packages for reinstall?
Use ghc-pkg to unregister libraries in /usr/local/lib. It is better not to
do that with libraries /usr/local/lib/ghc-7.6.3
Which installation options allow me to put all cabal packages into a
single folder?
AFAIK there are no such option
(not tutorial intended) Is it better to learn to solve the cabal hell?
I've heard about ghc-pkg
After the new constrains solver was introduced in cabal, the cabal hell
almost disappeared for me. The last cabal-install release introduced
sandboxes, and other improvements coming. I hope the cabal hell will di
There's no real magic bullet to this one.
Either the developers have to take some responsibility for ensuring changes
to shared assemblies don't impact other code that references those
assemblies or you really shouldn't be sharing them across projects. If the
things they're changing in these shared assemblies really happens that
often, you might consider migrating those parts or the methods they're
changing to each individual project. Sharing assemblies like this should
really only be done for absolute core functions that should change very,
very infrequently.
Ok, you created your own classloader and then loaded a class using it. The
question is - how does the thread classloader will know about that?
So, you must load the class using some classloader, and then set this
classloader as thread context classloader.
You could try to add the binary jar to your local eclipse plugins
directory.
You can find it here.
Very simply said, it absolutely will.
Enable USB debugging on your device.
Install phone drivers and connect your phone to your PC. When running your
application choose your phone from the list and run.
Read more about all steps on Using Hardware Devices, Android developer site
If you are using premium or above version of vs2012, try looking at
microsoft.fakes. it should cater to your situation.
Some reading:
I wonder why you need to store a single SqlConnection. It doesn't smell
great.
If you really do need to share a single SqlConnection across multiple
classes, dependency injection is likely a better option. Have a connection
factory instantiate a connection object and pass it around as required.
Otherwise, let the DBMS worry about controlling your connection resources.
Create, open and close a connection each time you need one.
I was able to make it work using RazorEngine by removing all @Html tags
from the partial view file.
string template = System.IO.File.ReadAllText(path);
string partialView = RazorEngine.Razor.Parse(template, model, "cachename");
After a long digging around I noticed that VS2013 come with a new addition;
SignalR - which as it turns out is associated with the ArteryFilter
problem.
So, in order to solve this problem, uncheck the "Enable Browser Link"
feature next to your Debug button and voila; filters works as expected
again. Still weird that VS2013 does not chain filters.
Also, note that this is a general ASP.NET feature and hence not limited to
MVC.
PRESERVED FOR HISTORY - ANSWER IS ABOVE
I am experiencing the same thing but so far it seems related to the new
IISExpress and not VS2013 perse. What worked fine in VS2012 suffers the
same fate as the one introduced by installing VS2013.
When executed through normal IIS the problem disappears, hence your code is
working fine. Let me know if you find a way to dis
Using the HttpContext object to pass a global object is a bad idea.
Normally, the business layer shouldn't even have a reference to System.Web.
Resharper has the "Extract Class from Parameters ..." refactoring to
automate what you describe. You can access it by right-clicking on your
method's name and choose Refactor > Extract > Extract Class from
Parameters. You can then select the parameters you want to include in the
class.
Resharper will update all references to the method to create and use the
new class.
In any case, more than 3 or 4 parameters is a code smell that suggests the
method may be trying to do too many things. When you cross that threshold
it's time to change your design and either consolidate the parameters or
split the method.
Have you included the System.Web assembly in the application?
using System.Web;
If not, try specifying the System.Web namespace, for example:
System.Web.HttpContext.Current
if (System.Environment.MachineName == "xxx")
Although, the preferred way of doing that is to have a base web.config
file, and creating XLST transforms for your different environments that
change the values based on how you publish your project. For example, we
have web.config, web.release.config, web.debug.config, and
web.staging.config. We have different values for caching, output caching,
the debug attribute, connection strings, email server settings, etc for
each environment and they all get swapped out at publish time.
Sub Application_Start(ByVal sender As Object, ByVal e As EventArgs)
' Fires when the application is started
Select Case LCase(System.Environment.MachineName)
Case "localhost", "development_domain"
DB_Environment = "DEVELOPMENT"
C
This is what I found on MSDN. I think this might help you.
The stream provider looks at the Content-Disposition header field and
determines an output Stream based on the presence of a filename parameter.
If a filename parameter is present in the Content-Disposition header field
then the body part is written to a FileStream, otherwise it is written to a
MemoryStream. This makes it convenient to process MIME Multipart HTML Form
data which is a combination of form data and file content.
You should check the Session for null like so:
public static bool BrowserSupportsJS
{
get
{
if(HttpContext.Current.Session == null)
return false;
return (HttpContext.Current.Session["js_support"] != null
&& ((bool)HttpContext.Current.Session["js_support"]));
}
}
I've needed use InputStream of Http Request. I have a WebApp and IOS App
that navigates to a aspx page, if the url request contains some parameters
i read the information in database and if i not find any parameters in url
request i read the request body and i work fine !
protected void Page_Load(object sender, EventArgs e)
{
try
{
if (string.IsNullOrEmpty(Request.QueryString["AdHoc"]) ==
false)
{
string v_AdHocParam = Request.QueryString["AdHoc"];
string [] v_ListParam = v_AdHocParam.Split(new
char[] {','});
if (v_ListParam.Length < 2)
{
DataContractJsonSerializer jsonSerializer = new
DataContractJsonSerializer
HttpContext.Current only gives you the context you want when you call it on
the thread that handles the incoming thread.
When calling it outside of such threads, you get null. That matches your
case, as Timer1_Elapsed is executed on a new thread.
You can get what you want like this:
string applicationPath = filterContext.HttpContext.Request.ApplicationPath;
string returnUrl =
filterContext.HttpContext.Request.RawUrl.Replace(applicationPath, "");
Did you recently upgrade from VS 2008 to VS 2012?
I think you still need to upgrade the web.config file then.
1) Remove this section in the Defin
Is my test flawed, or is there some web.config element I'm missing
here that would make HttpContext.Current resolve correctly after an
await?
Your test is not flawed and HttpContext.Current should not be null after
the await because in ASP.NET Web API when you await, this will ensure that
the code that follows this await is passed the correct HttpContext that was
present before the await.
Basically you have to create an ajax poll that calls a method that queries
the cache object. You cannot update a page from a cache expiration callback
as the thread that is used has nothing to do with a request object (it is
part of the asp.net/iis process and not anything to do with a HttpRequest
(which is where the .Context is populated)).
So basically on the page you need to display the data create either a
$.ajax or use an asp.net ATLAS timer (or straight .js ajax if you wish)
that calls either a asmx endpoint or a WebMethod which can query the
HttpContext.Current.Cache and return the appropriate
string/html/json/xml/whatever data.
UPDATE:
There are libraries available for the server updating clients without the
programmer understanding/implementing long polling/ polling techniques.
If you're on .NET 3.5 and up, you should check out the
System.DirectoryServices.AccountManagement (S.DS.AM) namespace. Read all
about it here:
Managing Directory Security Principals in the .NET Framework 3.5
MSDN docs on System.DirectoryServices.AccountManagement
Basically, you can define a domain context and easily find users and/or
groups in AD:
// set up domain context
using (PrincipalContext ctx = new PrincipalContext(ContextType.Domain))
{
// find a user
UserPrincipal user = UserPrincipal.FindByIdentity(ctx,
User.Identity.Name);
if(user != null)
{
Guid userGuid = user.Guid ?? Guid.Empty;
}
}
The new S.DS.AM makes it really easy to play around with users and groups
in AD!
What you are describing is really : duplicating the data into a second
(technically redundant) model, more suitable for query. If that is the
case, then sure : have fun with that - that isn't exactly uncommon.
However, before doing all that, you might want to try indexed views - it
could be that this solves most everything without you having to write all
the maintenance code.
I would suggest, however, not to "remove caching" - but simply "make the
cache expire at some point"; there's an important difference. Hitting the
database for the same data on every single request is not a great idea.
The cbxR1, cbxR2 should be the name of an input element, not the id.
<form ...>
<input type="checkbox" name="cbxR1" checked="checked" />
<input type="checkbox" name="cbxR2" checked="checked" />
</form>
MutateIncoming will be called on an NServiceBus worker thread and not on an
ASP.NET worker thread - hence no HTTP context.
Think of it - what would you expect the HTTP context to be when you're
handling an NServiceBus message?
If you need something from the user's session, you'd probably need to pass
some kind of session ID or correlation ID around, allowing you leave the
data in the right place when the reply message gets handled.
To do as you suggest:
For<RequestContext>().Use(ctx =>
{
//TODO: Create unittest.html as required.
SimpleWorkerRequest request = new SimpleWorkerRequest("/unittest",
@"c:inetpubwwwrootunittest", "unittest.html", null, new StringWriter());
HttpContext context = new HttpContext(request);
return HttpContext.Current = context;
});
As was recommended though, abstracting the dependency upon the context
would be a better long way to go.
You are getting this error because the object which you are trying to
iterate does not implement IEnumerable interface.
It seems like en.Value is an object here (I don't have VS with me to test
it right now). If it is then, you need to cast it into the appropriate type
and if your cast type is a collection and does not implement IEnumerable
interface, you need to call AsEnumerable() extension method by System.Linq
in place. Otherwise, as I said casting it into appropriate type should
work.
It gets the virtual path of the application root and makes it relative by
using the tilde (~) notation. It does not says anything about the
parameters of the path that you passed throught.
If you look at the begining of the returned string, it is the same. I'm
pretty sure that the URL in each request is different, with and without
LoginUser, so the complete returned string seems different base on that,
althought, for the function, they are the same. | http://www.w3hello.com/questions/-ASP-Net-hell-gt-HttpContext-Items- | CC-MAIN-2018-17 | refinedweb | 2,946 | 57.27 |
Hello,
I have been trying to figure out why my code won't compile. Any help and advice would be appreciated. Thank you!
I keep getting the following error message:
cannot find symbol -- method PrintIn(java.lang.String) with the following line highlighted:
System.out.printIn("Found " + i + " at " +
Here is the whole code:
public class BinarySearch { public static final int NOT_FOUND = -1; public static int binarySearch (Integer[] a, int x) { int low=0; int high = a.length - 1; int mid; while (low <=high) { mid = (low + high) / 2; if (a[mid].compareTo(x)<0) low = mid + 1; else if (a[mid].compareTo(x) > 0) high = mid - 1; else return mid; } return NOT_FOUND; } public static void main(String[] args) { int SIZE = 8; Integer[] a = new Integer[ SIZE ]; for (int i=0; i<SIZE; i++) a[i] = new Integer(i * 2); for (int i=0; i<SIZE*2; i++) System.out.printIn("Found " + i + " at " + binarySearch(a, new Integer(i))); } } | http://www.javaprogrammingforums.com/whats-wrong-my-code/30418-cannot-find-symbol-error-message.html | CC-MAIN-2014-10 | refinedweb | 158 | 65.42 |
Level of Difficulty: Beginner.
Are you having issues pip installing libraries to the right place? Could it be because you have more than one interpreter installed? Well… How many Python interpreters do you have installed?
The standard IDLE interpreter – the one that comes with the Python installation? The one that was installed when you were playing around with Visual Studios? No, no, no… It was VS Code! The one that was installed when you downloaded Anaconda?
Well… Now you no longer need to get bitten by that snake when untangling where you’re executing your Python code from.
Find the Python Interpreter Path (in 15 seconds or less)
Here’s what to do…
- Create a test script that will execute on the same interface you’re having issues with (PyCharm, Jupyter Notebook, IDLE, command prompt, Visual Studios, VS Code, etc.)
- Run the following code:
import sys print(sys.executable)
Now you know which interpreter to use to pip install your libraries.
Still having issues? Drop a comment below or reach out personally – jacqui.jm77@gmail.com
One thought on “[Python] Finding the Python Interpreter Path” | https://thejpanda.com/2020/01/28/python-finding-the-python-interpreter-path/ | CC-MAIN-2022-21 | refinedweb | 185 | 76.42 |
Intro CS using Python
This content was COPIED from BrainMass.com - View the original, and get the already-completed solution here!
Def y():
""" y does this function exist?
** just to illustrate a 0-input function...
"""
return 1
def w(x):
""" w computes thrice its input plus one
** plus it offers a chance to use the word "thrice"
input x: any number (int or float)
"""
return 3*(x+1)
def t(x):
""" t computes thrice its input plus one
** and shows how python is more precise than English
input x: any number (int or float)
"""
return 3*x + 1
def r(x,y):
""" r shows some less-common arithmetic operators
input x: any number (int or float)
input y: any number (int or float, more likely int)
"""
return ( x**2 % y ) + 2
def f(a,b):
""" f demonstrates the use of conditionals (if/elif/else)
** and that input parameters' names don't matter
input a: any number (int or float)
input b: any number (int or float)
"""
if a < b:
return (b-1) * (b-2)
else:
return (a+42) * (b+42)
PROBLEM
What is the return statement that satisfies these input/output constraints? Note: This function takes two inputs. You will need to pass them as (input1, input2)
an input of (3,4) returns 5
an input of (24,7) returns 25
an input of (42,1) returns 42.01190307520001
Use words to describe the solution steps, not just symbols.© BrainMass Inc. brainmass.com October 9, 2019, 6:39 pm ad1c9bdddf
Solution Preview
1. an input of (3,4) returns 5
There is no return statement to satisfy ... | https://brainmass.com/computer-science/python/intro-cs-using-python-93974 | CC-MAIN-2019-47 | refinedweb | 267 | 64.24 |
Survey period: 22 Apr 2013 to 29 Apr 2013
jQuery 2.0 has been released and drops support for IE8,7 and 6. Will you, too, drop IE6,7 and 8 support?
jkirkerx wrote:They tell me I'm crazy
public class Naerling : Lazy<Person>{
public void DoWork(){ throw new NotImplementedException(); }
}
Deveshdevil wrote:why should one compromise and not use JQuery...
AlexCode wrote:Most web users don't even know what a "browser" is.
AlexCode wrote:Don't think the results of this poll will actually be relevant to anything.
Will you move to jQuery 2.0 (and drop IE 6,7 and 8 support)?:We'll have to change according to the changes in technology.. | http://www.codeproject.com/script/Surveys/View.aspx?srvid=1441&msg=4546525 | CC-MAIN-2015-35 | refinedweb | 116 | 77.03 |
#include <wx/dataview.h>
wxDataViewTreeStore is a specialised wxDataViewModel for storing simple trees very much like wxTreeCtrl does and it offers a similar API.
This class actually stores the entire tree and the values (therefore its name) and implements all virtual methods from the base class so it can be used directly without having to derive any class from it, but it is mostly used from within wxDataViewTreeCtrl.
Notice that by default this class sorts all items with children before the leaf items. If this behaviour is inappropriate, you need to derive a custom class from this one and override either its HasDefaultCompare() method to return false, which would result in items being sorted just in the order in which they were added, or its Compare() function to compare the items using some other criterion, e.g. alphabetically.
Constructor.
Creates the invisible root node internally.
Destructor.
Append a container.
Append an item.
Delete all item in the model.
Delete all children of the item, but not the item itself.
Delete this item.
Return the number of children of item.
Returns the client data associated with the item.
Returns the icon to display in expanded containers.
Returns the icon of the item.
Returns the text of the item.
Returns the nth child item of item.
Inserts a container after previous.
Inserts an item after previous.
Inserts a container before the first child item or parent.
Inserts an item before the first child item or parent.
Sets the client data associated with the item.
Sets the expanded icon for the item.
Sets the icon for the item. | https://docs.wxwidgets.org/trunk/classwx_data_view_tree_store.html | CC-MAIN-2021-49 | refinedweb | 265 | 68.36 |
What is a good way to split a numpy array randomly into training and testing / validation dataset? Something similar to the cvpartition or crossvalind functions in Matlab.
If you want to divide the data set once in two halves, you can use
numpy.random.shuffle, or
numpy.random.permutation if you need to keep track of the indices:
import numpy # x is your dataset x = numpy.random.rand(100, 5) numpy.random.shuffle(x) training, test = x[:80,:], x[80:,:]
or
import numpy # x is your dataset x = numpy.random.rand(100, 5) indices = numpy.random.permutation(x.shape[0]) training_idx, test_idx = indices[:80], indices[80:] training, test = x[training_idx,:], x[test_idx,:]
There are many ways to repeatedly partition the same data set for cross validation. One strategy is to resample from the dataset, with repetition:
import numpy # x is your dataset x = numpy.random.rand(100, 5) training_idx = numpy.random.randint(x.shape[0], size=80) test_idx = numpy.random.randint(x.shape[0], size=20) training, test = x[training_idx,:], x[test_idx,:]
Finally, scikits.learn contains several cross validation methods (k-fold, leave-n-out, stratified-k-fold, ...). For the docs you might need to look at the examples or the latest git repository, but the code looks solid. | https://codedump.io/share/4dFQo2ez6XCV/1/numpy-how-to-splitpartition-a-dataset-array-into-training-and-test-datasets-for-eg-cross-validation | CC-MAIN-2017-04 | refinedweb | 209 | 60.51 |
Introduction: Java (Programming Language) for Beginners
This Instructable will show you the wonders of Java (programming language). You will also be able to DIY (Do It Yourself) at home. There is no cost involved within this Instructable. It's very easy, and requires no other programming language at all.
I have spread the main part over steps 3-5. Simply beacuse there is quite a lot of information.
Please rate this Instructable and leave comments, questions or statements. All questions, statements, and comments will be answered.
Step 1: What Is Java?
Java is just one of the hundreds of different programming languages in the world. Java language is an object-orientated programming language which was developed by Sun Microsystems. Java programmes are platform independent which means they can be run on any operating system with any type of processor as long as the Java interpreter is available on that system.
Step 2: What You Will Need
You will need the Java Software Development Kit from Sun's Java site. Follow the instructions on Sun's website to install it. Make sure that you add the java bin directory to your PATH environment variable. To find the Java Software Development Kit, go to the top right-hand corner of the screen and you will see a search bar. Type in: Java Software Development Kit. The the search results appear, find the one that says something along the lines of download.
Step 3: Writing Your First Java Programme:Part 1
You will need to write your Java programs using a text editor. When you type the examples that follow you must make sure that you use capital and small letters in the right places because Java is case sensitive. The first line you must type is:
public class Hello
This creates a class called Hello. All class names must start with a capital letter. The main part of the program must go between curly brackets after the class declaration. The curly brackets are used to group together everything inside them.
public class Hello
{
}
Step 4: Writing Your First Java Programme:Part 2
We must now create the main method which is the section that a program starts.
public class Hello
{
public static void main(String[] args)
{
}
}
You will see that the main method code has been moved over a few spaces from the left. This is called indentation and is used to make a program easier to read and understand.
Here is how you print the words Hello World on the screen:
public class Hello
{
public static void main(String[] args)
{
System.out.println("Hello World");
}
}
Step 5: Writing Your First Java Programme:Part 3
Make sure that you use a capital S in System because it is the name of a class. println is a method that prints the words that you put between the brackets after it on the screen. When you work with letters like in Hello World you must always put them between quotes. The semi-colon is used to show that it is the end of your line of code. You must put semi-colons after every line like this.
Step 6: Compiling the Programme
What we have just finished typing is called the source code. You must save the source code with the file name Hello.java before you can compile it. The file name must always be the same as the class name.
Make sure you have a command prompt open and then enter the following:
javac Hello.java
If you did everything right then you will see no errors messages and your program will be compiled. If you get errors then go through this lesson again and see where your mistake is.
Step 7: Running the Programme
Once your program has been compiled you will get a file called Hello.class. This is not like normal programs that you just type the name to run but it is actually a file containing the Java bytecode that is run by the Java interpreter. To run your program with the Java interpeter use the following command:
java Hello
Do not add .class on to the end of Hello. You will now see the following output on the screen:
Hello World
Congratulations! You have just made your first Java program.
11 Discussions
Your screenshots are of python not java code.
Thanks for that. I didn't pick up on that. I have fixed them up now.
um so I was wondering can u program like your own game with this Java stuff cause I'd like to do that I no Minecraft is based on Java: and was wondering?
I'd suggest you use Eclipse.
I'd suggest NetBeans mainly because by default it generates projects that can be easily compiled anywhere outside the IDE using Ant (instead of Eclipse's just calling the compiler directly) And the last 5 images need to be seriously rethought.
I have rethought the last 5 images and have changed them.
do I download i T for tablet,PC,,my phone,or does it matter?
Yea. I have the programme Eclipse already, but I wanted to do this a different way.
Does it matter if i download java 2 software development kit? are they the same thing?
This might be useful as notes for someone who has taken a class on Java, but as a stand alone work? Not useful. To many presumptions, no definitions, no why something works, no alternatives to make something work... This is 'how to write hello world', not 'java for beginners'.
I second your opinion. | http://www.instructables.com/id/Java-Programming-Language-For-Beginners/ | CC-MAIN-2018-30 | refinedweb | 929 | 73.68 |
29 May 2012 16:04 [Source: ICIS news]
LONDON (ICIS)--Polycarbonate (PC) producers in the European and the ?xml:namespace>
Bayer MaterialScience (BMS) and SABIC Innovative Plastics between them control over 80% of the PC market in Europe and the
“Niche commodities, such as PC, methyl di-p-phenylene isocyanate (MDI) and toluene di-isocyanate (TDI), have a concentrated supplier base and historically have been able to increase margins and generally maintain them even in a weak supply-driven environment,” said Bernstein in a note published today.
It claims BMS was able to achieve earnings growth straight after the 2008 financial crisis though rising raw material costs and very weak demand later in 2011 and in the first quarter of 2012 squeezed margins, especially for PC and MDI.
It added: “We believe consensus underestimates the power of this concentrated industry structure and its importance to attractive, sustainable margins for PC.”
The group also claims that earnings will be supported by a lack of new capacity coming onstream over the next couple of years.
“MaterialScience margins are set to improve through 2012 and beyond since the industry’s capacity surplus is working its way through the system and prices are increasing again,” it said.
Most new capacity came onstream in 2011 and early 2012, with the next wave not due onstream until 2014, Berstein said.
The analyst group today upgraded its target price for Bayer from €60 ($75) to €64 and boosted its outlook for the group to “outperform” from “market perform”. It also expects positive developments in Bayer’s CropScience and HealthCare divisions.
At 15:57 GMT, Bayer's share price was trading at €52.16 on the XETRA stock exchange, up 2.27% from the previous close.
BMS and SABIC Innovative Polymers did not respond immediately to a request for comments.
The graph below presents the ICIS price history for PC. It shows how PC prices recovered strongly following the 2008 financial crisis.
( | http://www.icis.com/Articles/2012/05/29/9565153/duopoly-market-structure-boosts-pc-margins-analyst.html | CC-MAIN-2014-42 | refinedweb | 324 | 50.46 |
The base64 module have functions which helps to encode the text or binary data into base64 format and decode the base64 data into text or binary data. The base64 module is used to encode and decode the data in the following ways:
Base64 Encoding
The base64 module provides the
b64encode() function. It encodes a bytes-like object using Base64 and returns the encoded bytes. Let's see how to use this function.
Note: Since we start with a string, we encode it first to a byte-like object using string.encode(). Later we convert it back to a string using string.decode(). This articles teaches more about the difference between byte objects and strings in Python.
import base64 data = "Python is a programming language" data_bytes = data.encode('ascii') base64_bytes = base64.b64encode(data_bytes) base64_string = base64_bytes.decode('ascii') print("Encoded Data: ", base64_string) # Output: Encoded Data: UHl0aG9uIGlzIGEgcHJvZ3JhbW1pbmcgbGFuZ3VhZ2U=
In the above example, first we convert the input string to byte-like objects and then encode those byte-like objects to base64 format.
Base64 Decoding
Decoding base64 string is opposite to that of encoding. The base64 module provides the
b64decode() function which decodes the Base64 encoded bytes-like object or ASCII string and returns the decoded bytes. Let's see how to use this function.
import base64 base64_string = "UHl0aG9uIGlzIGEgcHJvZ3JhbW1pbmcgbGFuZ3VhZ2U=" base64_bytes = base64_string.encode('ascii') data_bytes = base64.b64decode(base64_bytes) data = data_bytes.decode('ascii') print("Decoded Data:", data) # Output: Decoded Data: Python is a programming language
In the above example, first we convert the base64 strings into unencoded data bytes and then decode those bytes to get the original string.
Note: To prevent data corruption, make sure to use the same encoding format when converting from string to bytes, and from bytes to string.
Conclusion
In this tutorial, we have learned the basics of base64 encoding and decoding in Python. If you want to learn more about base64 encoding and decoding, you can visit the official documentation of the base64 module. | https://www.python-engineer.com/posts/base64-module/ | CC-MAIN-2022-40 | refinedweb | 323 | 56.96 |
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Re: Remote Shutdown using Java
From:
Eric Sosman <Eric.Sosman@sun.com>
Newsgroups:
comp.lang.java.programmer
Date:
Thu, 14 Jun 2007 17:29:13 -0400
Message-ID:
<1181856554.122861@news1nwk>
christopher_board@yahoo.co.uk wrote On 06/14/07 17:08,:
Hi all. I want to be able to shutdown remote computers using Java.
Below are the things that have been imported :
import java.awt.*;
import java.awt.event.ActionEvent;
import java.awt.event.ActionListener;
import java.rmi.*;
import java.rmi.server.*;
import javax.swing.*;
and below is the code
public void remoteShutdown_actionPerformed(ActionEvent e) {
try{
Runtime.getRuntime().exec("shutdown -m \\toshiba_cwb -s -t
99");
}catch(Exception ex){System.out.println("Fail to
ShutDown"+ex);}
The code works fine without the the -m and the computer name, however
when I put the -m \\computerName it won't shut the computer down but
no error messages have been displayed.
What is wrong with this.
Any help in this matter would be truly appreciated.
I have no idea what this "shutdown" program you're
using is; it certainly doesn't look like the one I know.
Nonetheless, I have a suspicion: are the two backslashes
part of the actual command syntax? That is, do you need
two backslashes in the executed command line? If so, be
aware that what you've written is only *one* backslash,
because of the way the Java compiler uses \ in strings
to introduce hard-to-type characters. To get two, you'll
need to double up each of them:
... ("shutdown -m \\\\toshiba_cwb ...");
Depending on what happens to the command line after
you launch it, even that might not be enough. For example,
if the \ is also special to the command processor ("shell"),
then you may need to double it yet again or escape it by
whatever mechanism the shell uses:
... ("shutdown -m \\\\\\\\toshiba_cwb ...");
(Intepretation: The Java compiler generates one "delivered"
backslash for each pair in the source, making four. Then
the shell makes one backslash out of each pair that *it*
sees, making two. YMMV.)
--
Eric.Sosman@sun.com
Generated by PreciseInfo ™
"The Jews are the master robbers of the modern age."
-- Napoleon Bonaparte | http://preciseinfo.org/Convert/Articles_Java/Listener_Experts/Java-Listener-Experts-070615002913.html | CC-MAIN-2021-49 | refinedweb | 370 | 58.89 |
Okay, I am very new to ruby. I have written several standard ruby programs and now I am starting to learn ROR. I am a bit confused though. In a standard ruby program I create a method and then in the main program I pass a variable to the method and have the method print out a value. Now, I know I use the controller file and the index file to do this, but I am unsure exactly how to include the method and to call it.
def print(value)
puts(value)
end
print(10)
I need to do something like this in ROR, but I am not sure just how. It is simple in just a standard ruby program. Any help would be appreciated. | http://forums.devshed.com/ruby-programming-142/ror-confusion-using-methods-651849.html | CC-MAIN-2016-30 | refinedweb | 125 | 81.33 |
Function annotations are a Python 3 feature that lets you add arbitrary metadata to function arguments and return value. They were part of the original Python 3.0 spec.
In this tutorial I’ll show you how to take advantage of general-purpose function annotations and combine them with decorators. You will also learn about the pros and cons of function annotations, when it is appropriate to use them, and when it is best to use other mechanisms like docstrings and plain decorators.
Function Annotations
Function annotations are specified in PEP-3107. The main motivation was to provide a standard way to associate metadata to function arguments and return value. A lot of community members found novel use cases, but used different methods such as custom decorators, custom docstring formats and adding custom attributes to the function object.
It’s important to understand that Python doesn’t bless the annotations with any semantics. It purely provides a nice syntactic support for associating metadata as well as an easy way to access it. Also, annotations are totally optional.
Let’s take a look at an example. Here is a function foo() that takes three arguments called a, b and c and prints their sum. Note that foo() returns nothing. The first argument a is not annotated. The second argument b is annotated with the string ‘annotating b’, and the third argument c is annotated with type int. The return value is annotated with the type float. Note the “->” syntax for annotating the return value.
python
def foo(a, b: 'annotating b', c: int) -> float:
print(a + b + c)
The annotations have no impact whatsoever on the execution of the function. Let’s call foo() twice: once with int arguments and once with string arguments. In both cases, foo() does the right thing, and the annotations are simply ignored.
```python foo(‘Hello’, ‘, ‘, ‘World!’) Hello, World!
foo(1, 2, 3) 6 ```
Default Arguments
Default arguments are specified after the annotation:
```python def foo(x: ‘an argument that defaults to 5’ = 5): print(x)
foo(7) 7
foo() 5 ```
Accessing Function Annotations
The function object has an attribute called ‘annotations’. It is a mapping that maps each argument name to its annotation. The return value annotation is mapped to the key ‘return’, which can’t conflict with any argument name because ‘return’ is a reserved word that can’t serve as an argument name. Note that it is possible to pass a keyword argument named return to a function:
```python def bar(*args, **kwargs: ‘the keyword arguments dict’): print(kwargs[‘return’])
d = {‘return’: 4} bar(**d) 4 ```
Let’s go back to our first example and check its annotations:
```python def foo(a, b: ‘annotating b’, c: int) -> float: print(a + b + c)
print(foo.annotations) {‘c’: <class ‘int’>, ‘b’: ‘annotating b’, ‘return’: <class ‘float’>} ```
This is pretty straightforward. If you annotate a function with an arguments array and/or keyword arguments array, then obviously you can’t annotate individual arguments.
```python def foo(*args: ‘list of unnamed arguments’, **kwargs: ‘dict of named arguments’): print(args, kwargs)
print(foo.annotations) {‘args’: ‘list of unnamed arguments’, ‘kwargs’: ‘dict of named arguments’} ```
If you read the section about accessing function annotations in PEP-3107, it says that you access them through the ‘func_annotations’ attribute of the function object. This is out of date as of Python 3.2. Don’t be confused. It’s simply the ‘annotations’ attribute.
What Can You Do With Annotations?
This is the big question. Annotations have no standard meaning or semantics. There are several categories of generic uses. You can use them as better documentation and move argument and return value documentation out of the docstring. For example, this function:
python
def div(a, b):
"""Divide a by b
args:
a - the dividend
b - the divisor (must be different than 0)
return:
the result of dividing a by b
"""
return a / b
Can be converted to:
python
def div(a: 'the dividend',
b: 'the divisor (must be different than 0)') -> 'the result of dividing a by b':
"""Divide a by b"""
return a / b
While the same information is captured, there are several benefits to the annotations version:
- If you rename an argument, the documentation docstring version may be out of date.
- It is easier to see if an argument is not documented.
- There is no need to come up with a special format of argument documentation inside the docstring to be parsed by tools. The annotations attribute provides a direct, standard mechanism of access.
Another usage that we will talk about later is optional typing. Python is dynamically typed, which means you can pass any object as an argument of a function. But often functions will require arguments to be of a specific type. With annotations you can specify the type right next to the argument in a very natural way.
Remember that just specifying the type will not enforce it, and additional work (a lot of work) will be needed. Still, even just specifying the type can make the intent more readable than specifying the type in the docstring, and it can help users understand how to call the function.
Yet another benefit of annotations over docstring is that you can attach different types of metadata as tuples or dicts. Again, you can do that with docstring too, but it will be text-based and will require special parsing.
Finally, you can attach a lot of metadata that will be used by special external tools or at runtime via decorators. I’ll explore this option in the next section.
Multiple Annotations
Suppose you want to annotate an argument with both its type and a help string. This is very easy with annotations. You can simply annotate the argument with a dict that has two keys: ‘type’ and ‘help’.
```python def div(a: dict(type=float, help=’the dividend’), b: dict(type=float, help=’the divisor (must be different than 0)’) ) -> dict(type=float, help=’the result of dividing a by b’): “"”Divide a by b””” return a / b
print(div.annotations) {‘a’: {‘help’: ‘the dividend’, ‘type’: float}, ‘b’: {‘help’: ‘the divisor (must be different than 0)’, ‘type’: float}, ‘return’: {‘help’: ‘the result of dividing a by b’, ‘type’: float}} ```
Combining Python Annotations and Decorators
Annotations and decorators go hand in hand. For a good introduction to Python decorators, check out my two tutorials: Deep Dive Into Python Decorators and Write Your Own Python Decorators.
First, annotations can be fully implemented as decorators. You can just define an @annotate decorator and have it take an argument name and a Python expression as arguments and then store them in the target function’s annotations attribute. This can be done for Python 2 as well.
However, the real power of decorators is that they can act on the annotations. This requires coordination, of course, about the semantics of annotations.
Let’s look at an example. Suppose we want to verify that arguments are in a certain range. The annotation will be a tuple with the minimum and maximum value for each argument. Then we need a decorator that will check the annotation of each keyword argument, verify that the value is within the range, and raise an exception otherwise. Let’s start with the decorator:
python
def check_range(f):
def decorated(*args, **kwargs):
for name, range in f.__annotations__.items():
min_value, max_value = range
if not (min_value <= kwargs[name] <= max_value):
msg = 'argument {} is out of range [{} - {}]'
raise ValueError(msg.format(name, min_value, max_value))
return f(*args, **kwargs)
return decorated
Now, let’s define our function and decorate it with the @check_range decorators.
python
@check_range
def foo(a: (0, 8), b: (5, 9), c: (10, 20)):
return a * b - c
Let’s call foo() with different arguments and see what happens. When all arguments are within their range, there is no problem.
python
foo(a=4, b=6, c=15)
9
But if we set c to 100 (outside of the (10, 20) range) then an exception is raised:
python
foo(a=4, b=6, c=100)
ValueError: argument c is out of range [10 - 20]
When Should You Use Decorators Instead of Annotations?
There are several situations where decorators are better than annotations for attaching metadata.
One obvious case is if your code needs to be compatible with Python 2.
Another case is if you have a lot of a metadata. As you saw earlier, while it’s possible to attach any amount of metadata by using dicts as annotations, it is pretty cumbersome and actually hurts readability.
Finally, if the metadata is supposed to be operated on by a specific decorator, it may be better to associate the metadata as arguments for the decorator itself.
Dynamic Annotations
Annotations are just a dict attribute of a function.
python
type(foo.__annotations__)
dict
This means you can modify them on the fly while the program is running. What are some use cases? Suppose you want to find out if a default value of an argument is ever used. Whenever the function is called with the default value, you can increment the value of an annotation. Or maybe you want to sum up all the return values. The dynamic aspect can be done inside the function itself or by a decorator.
```python def add(a, b) -> 0: result = a + b add.annotations[‘return’] += result return result
print(add.annotations[‘return’]) 0
add(3, 4) 7 print(add.annotations[‘return’]) 7
add(5, 5) 10 print(add.annotations[‘return’]) 17 ```
Conclusion
Function annotations are versatile and exciting. They have the potential to usher in a new era of introspective tools that help developers master more and more complex systems. They also offer the more advanced developer a standard and readable way to associate metadata directly with arguments and return value in order to create custom tools and interact with decorators. But it takes some work to benefit from them and utilize their potential.<< | https://code.tutsplus.com/tutorials/python-3-function-annotations--cms-25689 | CC-MAIN-2019-18 | refinedweb | 1,657 | 54.22 |
I'm reading a file in my C program and comparing every word in it with my word, which is entered via command line argument. But I get crashes, and I can't understand what's wrong. How do I track such errors? What is wrong in my case?
My compiler is clang. The code compiles fine. When running it says 'segmentation fault'.
Here is the code.
#include <stdio.h>
#include <string.h>
int main(int argc, char* argv[])
{
char* temp = argv[1];
char* word = strcat(temp, "\n");
char* c = "abc";
FILE *input = fopen("/usr/share/dict/words", "r");
while (strcmp(word, c))
{
char* duh = fgets(c, 20, input);
printf("%s", duh);
}
if (!strcmp (word, c))
{
printf("FOUND IT!\n");
printf("%s\n%s", word, c);
}
fclose(input);
}
The issue here is that you are trying to treat strings in C as you might in another language (like C++ or Java), in which they are resizable vectors that you can easily append or read an arbitrary amount of data into.
C strings are much lower level. They are simply an array of characters (or a pointer to such an array; arrays can be treated like pointers to their first element in C anyhow), and the string is treated as all of the characters within that array up to the first null character. These arrays are fixed size; if you want a string of an arbitrary size, you need to allocate it yourself using
malloc(), or allocate it on the stack with the size that you would like.
One thing here that is a little confusing is you are using a non-standard type
string. Given the context, I'm assuming that's coming from your
cs50.h, and is just a typedef to
char *. It will probably reduce confusion if you actually use
char * instead of
string; using a typedef obscures what's really going on.
Let's start with the first problem.
string word = strcat(argv[1], "\n");
strcat() appends the second string onto the first; it starts from the null terminator of the first string, and replaces that with the first character of the second string, and so on, until it reaches a null in the second string. In order for this to work, the buffer containing the first string needs to have enough room to fit the second one. If it does not, you may overwrite arbitrary other memory, which could cause your program to crash or have all kinds of other unexpected behavior.
Here's an illustration. Let's say that
argv[1] contains the word
hello, and the buffer has exactly as much space as it needs for this. After it is some other data; I've filled in
other for the sake of example, though it won't actually be that, it could be anything, and it may or may not be important:
+---+---+---+---+---+---+---+---+---+---+---+---+ | h | e | l | l | o | \0| o | t | h | e | r | \0| +---+---+---+---+---+---+---+---+---+---+---+---+
Now if you use
strcat() to append
"\n", you will get:
+---+---+---+---+---+---+---+---+---+---+---+---+ | h | e | l | l | o | \n| \0| t | h | e | r | \0| +---+---+---+---+---+---+---+---+---+---+---+---+
You can see that we've overwritten the
other data that was after
hello. This may cause all kinds of problems. To fix this, you need to copy your
argv[1] into a new string, that has enough room for it plus one more character (and don't forget the trailing null). You can call
strlen() to get the length of the string, then add 1 for the
\n, and one for the trailing null, to get the length that you need.
Actually, instead of trying to add a
\n to the word you get in from the command line, I would recommend stripping off the
\n from your input words, or using
strncmp() to compare all but the last character (the
\n). In general, it's best in C to avoid appending strings, as appending strings means you need to allocate memory and copy things around, and it can be easy to make mistakes doing so, as well as being inefficient. Higher level languages usually take care of the details for you, making it easier to append strings, though still just as inefficient.
After your edit, you changed this to:
char* temp = argv[1]; char* word = strcat(temp, "\n");
However, this has the same problem. A
char * is a pointer to a character array. Your
temp variable is just copying the pointer, not the actual value; it is still pointing to the same buffer. Here's an illustration; I'm making up addresses for the purposes of demonstration, in the real machine there will be more objects in between these things, but this should suffice for the purpose of demonstration.
+------------+---------+-------+ | name | address | value | +------------+---------+-------+ | argv | 1000 | 1004 |-------+ | argv[0] | 1004 | 1008 | --+ <-+ | argv[1] | 1006 | 1016 | --|---+ | argv[0][0] | 1008 | 'm' | <-+ | | argv[0][1] | 1009 | 'y' | | | argv[0][2] | 1010 | 'p' | | | argv[0][3] | 1011 | 'r' | | | argv[0][4] | 1012 | 'o' | | | argv[0][5] | 1013 | 'g' | | | argv[0][6] | 1014 | 0 | | | argv[1][0] | 1016 | 'w' | <-+ <-+ | argv[1][1] | 1017 | 'o' | | | argv[1][2] | 1018 | 'r' | | | argv[1][3] | 1019 | 'd' | | | argv[1][4] | 1020 | 0 | | +------------+---------+-------+ |
Now when you create your
temp variable, all you are doing is copying
argv[1] into a new
char *:
+------------+---------+-------+ | | name | address | value | | +------------+---------+-------+ | | temp | 1024 | 1016 | --+ +------------+---------+-------+
As a side note, you also shouldn't ever try to access
argv[1] without checking that
argc is greater than 1. If someone doesn't pass any arguments in, then
argv[1] itself is invalid to access.
I'll move on to the next problem.
string c = "abc"; // ... char* duh = fgets(c, 20, input);
Here, you are referring to the static string
"abc". A string that appears literally in the source, like
"abc", goes into a special, read-only part of the memory of the program. Remember what I said;
string here is just a way of saying
char *. So
c is actually just a pointer into this read-only section of memory; and it has only enough room to store the characters that you provided in the text (4, for
abc and the null character terminating the string).
fgets() takes as its first argument a place to store the string that it is reading, and its second the amount of space that it has. So you are trying to read up to 20 bytes, into a read-only buffer that only has room for 4.
You need to either allocate space for reading on the stack, using, for example:
char c[20];
Or dynamically, using
malloc():
char *c = malloc(20); | https://codedump.io/share/Uu42HLkuo8pe/1/reading-file-in-c | CC-MAIN-2016-44 | refinedweb | 1,096 | 74.93 |
If you’ve been keeping up with my development content, you’ll remember that I recently wrote Build Your First .NET Core Application with MongoDB Atlas, which focused on building a console application that integrated with MongoDB. While there is a fit for MongoDB in console applications, many developers are going to find it more valuable in web applications.
In this tutorial, we’re going to expand upon the previous and create a RESTful API with endpoints that perform basic create, read, update, and delete (CRUD) operations against MongoDB Atlas.
To be successful with this tutorial, you’ll need to have a few things taken care of first:
We won’t go through the steps of deploying a MongoDB Atlas cluster or configuring it with user and network rules. If this is something you need help with, check out a previous tutorial that was published on the topic.
We’ll be using .NET Core 6.0 in this tutorial, but other versions may still work. Just take the version into consideration before continuing.
To kick things off, we’re going to create a fresh .NET Core project using the web application template that Microsoft offers. To do this, execute the following commands from the CLI:
dotnet new webapi -o MongoExample cd MongoExample dotnet add package MongoDB.Driver
The above commands will create a new web application project for .NET Core and install the latest MongoDB driver. We’ll be left with some boilerplate files as part of the template, but we can remove them.
Inside the project, delete any file related to
WeatherForecast and similar.
Before we start designing each of the RESTful API endpoints with .NET Core, we need to create and configure our MongoDB service and define the data model for our API.
We’ll start by working on our MongoDB service, which will be responsible for establishing our connection and directly working with documents within MongoDB. Within the project, create “Models/MongoDBSettings.cs” and add the following C# code:
namespace MongoExample.Models; public class MongoDBSettings { public string ConnectionURI { get; set; } = null!; public string DatabaseName { get; set; } = null!; public string CollectionName { get; set; } = null!; }
The above
MongoDBSettings class will hold information about our connection, the database name, and the collection name. The data we plan to store in these class fields will be found in the project’s “appsettings.json” file. Open it and add the following:
{ "Logging": { "LogLevel": { "Default": "Information", "Microsoft.AspNetCore": "Warning" } }, "AllowedHosts": "*", "MongoDB": { "ConnectionURI": "ATLAS_URI_HERE", "DatabaseName": "sample_mflix", "CollectionName": "playlist" } }
Specifically take note of the
MongoDB field. Just like with the previous example project, we’ll be using the “sample_mflix” database and the “playlist” collection. You’ll need to grab the
ConnectionURI string from your MongoDB Atlas Dashboard.
With the settings in place, we can move onto creating the service.
Create “Services/MongoDBService.cs” within your project and add the following:
using MongoExample.Models; using Microsoft.Extensions.Options; using MongoDB.Driver; using MongoDB.Bson; namespace MongoExample.Services; public class MongoDBService { private readonly IMongoCollection<Playlist> _playlistCollection; public MongoDBService(IOptions<MongoDBSettings> mongoDBSettings) { MongoClient client = new MongoClient(mongoDBSettings.Value.ConnectionURI); IMongoDatabase database = client.GetDatabase(mongoDBSettings.Value.DatabaseName); _playlistCollection = database.GetCollection<Playlist>(mongoDBSettings.Value.CollectionName); } public async Task<List<Playlist>> GetAsync() { } public async Task CreateAsync(Playlist playlist) { } public async Task AddToPlaylistAsync(string id, string movieId) {} public async Task DeleteAsync(string id) { } }
In the above code, each of the asynchronous functions were left blank on purpose. We’ll be populating those functions as we create our endpoints. Instead, make note of the constructor method and how we’re taking the passed settings that we saw in our “appsettings.json” file and setting them to variables. In the end, the only variable we’ll ever interact with for this example is the
_playlistCollection variable.
With the service available, we need to connect it to the application. Open the project’s “Program.cs” file and add the following at the top:
using MongoExample.Models; using MongoExample.Services; var builder = WebApplication.CreateBuilder(args); builder.Services.Configure<MongoDBSettings>(builder.Configuration.GetSection("MongoDB")); builder.Services.AddSingleton<MongoDBService>();
You’ll likely already have the
builder variable in your code because it was part of the boilerplate project, so don’t add it twice. What you’ll need to add near the top is an import to your custom models and services as well as configuring the service.
Remember the
MongoDB field in the “appsettings.json” file? That is the section that the
GetSection function is pulling from. That information is passed into the singleton service that we created.
With the service created and working, with the exception of the incomplete asynchronous functions, we can focus on creating a data model for our collection.
Create “Models/Playlist.cs” and add the following C# code:
using MongoDB.Bson; using MongoDB.Bson.Serialization.Attributes; using System.Text.Json.Serialization; namespace MongoExample.Models; public class Playlist { [BsonId] [BsonRepresentation(BsonType.ObjectId)] public string? Id { get; set; } public string username { get; set; } = null!; [BsonElement("items")] [JsonPropertyName("items")] public List<string> movieIds { get; set; } = null!; }
There are a few things happening in the above class that take it from a standard C# class to something that can integrate seamlessly into a MongoDB document.
First, you might notice the following:
[BsonId] [BsonRepresentation(BsonType.ObjectId)] public string? Id { get; set; }
We’re saying that the
Id field is to be represented as an ObjectId in BSON and the
_id field within MongoDB. However, when we work with it locally in our application, it will be a string.
The next thing you’ll notice is the following:
[BsonElement("items")] [JsonPropertyName("items")] public List<string> movieIds { get; set; } = null!;
Even though we plan to work with
movieIds within our C# application, in MongoDB, the field will be known as
items and when sending or receiving JSON, the field will also be known as
items instead of
movieIds.
You don’t need to define custom mappings if you plan to have your local class field match the document field directly. Take the
username field in our example. It has no custom mappings, so it will be
username in C#,
username in JSON, and
username in MongoDB.
Just like that, we have a MongoDB service and document model for our collection to work with for .NET Core.
When building CRUD endpoints for this project, we’ll need to bounce between two different locations within our project. We’ll need to define the endpoint within a controller and do the work within our service.
Create “Controllers/PlaylistController.cs” and add the following code:
using System; using Microsoft.AspNetCore.Mvc; using MongoExample.Services; using MongoExample.Models; namespace MongoExample.Controllers; [Controller] [Route("api/[controller]")] public class PlaylistController: Controller { private readonly MongoDBService _mongoDBService; public PlaylistController(MongoDBService mongoDBService) { _mongoDBService = mongoDBService; } [HttpGet] public async Task<List<Playlist>> Get() {} [HttpPost] public async Task<IActionResult> Post([FromBody] Playlist playlist) {} [HttpPut("{id}")] public async Task<IActionResult> AddToPlaylist(string id, [FromBody] string movieId) {} [HttpDelete("{id}")] public async Task<IActionResult> Delete(string id) {} }
In the above
PlaylistController class, we have a constructor method that gains access to our singleton service class. Then we have a series of endpoints for this particular controller. We could add far more endpoints than this to our controller, but it’s not necessary for this example.
Let’s start with creating data through the POST endpoint. To do this, it’s best to start in the “Services/MongoDBService.cs” file:
public async Task CreateAsync(Playlist playlist) { await _playlistCollection.InsertOneAsync(playlist); return; }
We had set the
_playlistCollection in the constructor method of the service, so we can now use the
InsertOneAsync method, taking a passed
Playlist variable and inserting it. Jumping back into the “Controllers/PlaylistController.cs,” we can add the following:
[HttpPost] public async Task<IActionResult> Post([FromBody] Playlist playlist) { await _mongoDBService.CreateAsync(playlist); return CreatedAtAction(nameof(Get), new { id = playlist.Id }, playlist); }
What we’re saying is that when the endpoint is executed, we take the
Playlist object from the request, something that .NET Core parses for us, and pass it to the
CreateAsync function that we saw in the service. After the insert, we return some information about the interaction.
It’s important to note that in this example project, we won’t be validating any data flowing from HTTP requests.
Let’s jump to the read operations.
Head back into the “Services/MongoDBService.cs” file and add the following function:
public async Task<List<Playlist>> GetAsync() { return await _playlistCollection.Find(new BsonDocument()).ToListAsync(); }
The above
Find operation will return all documents that exist in the collection. If you wanted to, you could make use of the
FindOne or provide filter criteria to return only the data that you want. We’ll explore filters shortly.
With the service function ready, add the following endpoint to the “Controllers/PlaylistController.cs” file:
[HttpGet] public async Task<List<Playlist>> Get() { return await _mongoDBService.GetAsync(); }
Not so bad, right? We’ll be doing the same thing for the other endpoints, more or less.
The next CRUD stage to take care of is the updating of data. Within the “Services/MongoDBService.cs” file, add the following function:
public async Task AddToPlaylistAsync(string id, string movieId) { FilterDefinition<Playlist> filter = Builders<Playlist>.Filter.Eq("Id", id); UpdateDefinition<Playlist> update = Builders<Playlist>.Update.AddToSet<string>("movieIds", movieId); await _playlistCollection.UpdateOneAsync(filter, update); return; }
Rather than making changes to the entire document, we’re planning on adding an item to our playlist and nothing more. To do this, we set up a match filter to determine which document or documents should receive the update. In this case, we’re matching on the id which is going to be unique. Next, we’re defining the update criteria, which is an
AddToSet operation that will only add an item to the array if it doesn’t already exist in the array.
The
UpdateOneAsync method will only update one document even if the match filter returned more than one match.
In the “Controllers/PlaylistController.cs” file, add the following endpoint to pair with the
AddToPlayListAsync function:
[HttpPut("{id}")] public async Task<IActionResult> AddToPlaylist(string id, [FromBody] string movieId) { await _mongoDBService.AddToPlaylistAsync(id, movieId); return NoContent(); }
In the above PUT endpoint, we are taking the
id from the route parameters and the
movieId from the request body and using them with the
AddToPlaylistAsync function.
This brings us to our final part of the CRUD spectrum. We’re going to handle deleting of data.
In the “Services/MongoDBService.cs” file, add the following function:
public async Task DeleteAsync(string id) { FilterDefinition<Playlist> filter = Builders<Playlist>.Filter.Eq("Id", id); await _playlistCollection.DeleteOneAsync(filter); return; }
The above function will delete a single document based on the filter criteria. The filter criteria, in this circumstance, is a match on the id which is always going to be unique. Your filters could be more extravagant if you wanted.
To bring it to an end, the endpoint for this function would look like the following in the “Controllers/PlaylistController.cs” file:
[HttpDelete("{id}")] public async Task<IActionResult> Delete(string id) { await _mongoDBService.DeleteAsync(id); return NoContent(); }
We only created four endpoints, but you could take everything we did and create 100 more if you wanted to. They would all use a similar strategy and can leverage everything that MongoDB has to offer.
You just saw how to create a simple four endpoint RESTful API using .NET Core and MongoDB. This was an expansion to the previous tutorial, which went over the same usage of MongoDB, but in a console application format rather than web application.
Like I mentioned, you can take the same strategy used here and apply it towards more endpoints, each doing something critical for your web application.
Got a question about the driver for .NET? Swing by the MongoDB Community Forums!
A video version of this tutorial can be found below.
This content first appeared on MongoDB. | https://www.thepolyglotdeveloper.com/2022/02/create-restful-api-dotnet-core-mongodb/ | CC-MAIN-2022-40 | refinedweb | 1,968 | 50.02 |
Provided by: manpages-dev_5.02-1_all
NAME
getpriority, setpriority - get/set program scheduling priority
SYNOPSIS
#include <sys/time.h> #include <sys/resource.h> int getpriority(int which, id_t who); int setpriority(int which, id_t who, int prio);
DESCRIPTION
The it afterward to determine if -1 is an error or a legitimate value. setpriority() returns 0 on success. On error, it returns -1 and sets errno to indicate the cause of the error. TO
POSIX.1-2001, POSIX.1-2008, SVr4, 4.4BSD (these interfaces first appeared in 4.2BSD).
NOTES
For further details on the nice value, see sched(7). Note: the addition of the "autogroup" feature in Linux 2.6.38 means that the nice value no longer has its traditional effect in many circumstances. For details, see sched(7). A child created by fork(2) inherits its parent's nice value. The nice value is preserved across execve(2). and later. Including <sys/time.h> is not required these days, but increases portability. (Indeed, <sys/resource.h> defines the rusage structure with fields of type struct timeval defined in <sys/time.h>.) C library/kernel differences Within the kernel, nice values are actually represented using the. (Thus, the kernel's 40..1 range corresponds to the range -20..19 as seen by user space.)
BUGS
According ALSO
nice(1), renice(1), fork(2), capabilities(7), sched(7) Documentation/scheduler/sched-nice-design.txt in the Linux kernel source tree (since Linux 2.6.23)
COLOPHON
This page is part of release 5.02 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/eoan/man2/setpriority.2.html | CC-MAIN-2019-47 | refinedweb | 281 | 53.47 |
Why Scala Doesn’t Have Java Interfaces (And Doesn’t Need Them)
Why Scala Doesn’t Have Java Interfaces (And Doesn’t Need Them)
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One thing that puzzled me a little when I started out in Scala was the lack of interfaces and the choice of “traits” instead, which seemed to be a hybrid between interfaces and abstract classes from Java. How are you supposed to achieve clean polymorphism without interfaces I wondered?
Well, turns out what has a single solution in Java actually has a multitude of solutions in Scala, the obvious one is to use traits as straight interface replacements, and that certainly works.
However, there are two even more elegant solutions available in Scala that doesn’t force you to implement/extend a specific trait:Function Passing
Given Scala’s functional aspects, one obvious alternative to traits is function passing. Say you want to do something with an InputStream, and you want to force the client of a function to have an “interface” that takes an InputStream as an argument and returns just about anything (or nothing). Well, that’s easy-peasy with Scala:
def getAbsoluteResource(path: String)(op: InputStream => Any) = {..}In the actual call, we may replace the “someInputStreamFunction” with just about any other function, from any other object that conforms to the same contract. Why should we be forced extend specific traits and have specific names for our functions if that just works?
def someInputStreamFunction(io: InputStream): Unit = {..}
// call the function:
getAbsoluteResource("/myPath.properties")(someInputStreamFunction)
Also, this style of programming allows for some interesting things such as “around blocks”, which are in fact functions that take functions/function closures as an argument. This certainly makes lifecycle management a lot easier than the customary verbose Java-style of defining lifecycle methods on interfaces.
An example of this sort of “around” pattern with function passing is the below example of JPA transaction management with the recursivity-jpa library:
transaction{Structural Types/”Static Duck Typing”
..your code within the transaction/persistence context goes here..
}
How many times have you created interfaces like “Identifiable” with a “getId” method in Java merely to deal with entities that need persisting in a more generic way? I have, a lot. Well, Scala has a powerful little feature called “Structural Types” that can get rid of all of this - it is basically a means of enforcing a contract on a type without forcing it to extend anything else. You do this as follows:
def someFunction(identifiable: {def id: Long}){Here we have basically instructed the compiler to say that the function “someFunction” can take any argument that is of type “AnyRef”, but the AnyRef argument MUST have a function on it called “id” that returns a Long. Incidentally, for case classes, “def” will work fine with val’s and var’s defined in the case class constructor. Pretty neat, huh? Of course, you can further use structural types with casts in “asInstanceOf” or “isInstanceOf”, or combine it with Generics. In other words, you can enforce an old-school “Java interface” without having to explicitly implement an interface. The absolute prime example where this is useful is the case I already mentioned: in persistence code where entities need to be identifiable, but you don’t want to force a trait/interface down the throats of the entities.
..do stuff..
}
So there you have it, a small primer of why Scala doesn’t need Java interfaces: Scala simply has more elegant ways of dealing with the same problem, without any of the weight that comes with Java interfaces.
This may be pretty basic stuff for the seasoned Scala developer, but I thought I’d put it down in writing so newbies have an easily accessible explanation for an issue that is quite a common question that I come across with people that are new to Scala.
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When you create an application which contains a tabbed view and define a manifest file (as explained on this MSDN page), you would expect the tabbed view to display with the active theme. But it does not! Why?
System.Windows.Forms.TabPage does not support the XP visual style, as shown in the screen shot (left tabbed view), but
System.Windows.Forms.TabControl does.
I searched the web, hoping to find a solution. I only found people who observed exactly the same problem as I did; and there was no solution readily available to C# programmers.
Trying to set the
TabPage.BackColor to transparent did not help. I had to implement my own
OnPaintBackground method and draw the themed background myself, if I wanted the tabbed view to display just like those found in non-.NET applications (standard applications don't have a
TabPage control above their
TabControl; the
TabPage is mainly used to simplify the manipulation of the tab page contents).
In order to draw the themed background, you have to call
uxtheme.dll function
DrawThemeBackground, which means calling into unmanaged code. I hacked the sources of Microsoft's sample ThemeExplorer in order to extract the C++ wrappers needed to call
uxtheme.dll. I implemented the following functions in an unmanaged C++ DLL (that is, pure Win32 code) named
OPaC.uxTheme.Win32.dll :
bool Wrapper_IsAppThemed ();
trueon XP if the user has activated the "Windows XP style" in the display properties. On systems which do not support
uxtheme.dll, the function returns
false.
bool Wrapper_DrawBackground (const wchar_t* name, const wchar_t* part, const wchar_t* state, HDC hdc, ...);
"BUTTON"/
"CHECKBOX"and
"TAB"/
"BODY"would be valid names/parts, for instance).
"UNCHECKEDNORMAL",
"UNCHECKEDHOT",
"UNCHECKEDPRESSED",
"UNCHECKEDDISABLED",
"CHECKEDNORMAL", etc.)
bool Wrapper_DrawThemeParentBackground (HWND hwnd, HDC hdc);
bool Wrapper_DrawThemeParentBackgroundRect (HWND hwnd, HDC hdc, int ox, int oy, int dx, int dy);
bool Wrapper_GetTextColor (const wchar_t* name, const wchar_t* part, const wchar_t* state, int* r, int* g, int* b);
Wrapper_DrawBackground.
All these functions degrade gracefully if there is no
uxtheme.dll present on the system (they just return
false ). They are accessible from .NET as static functions published by the
OPaC.uxTheme.Wrapper class. The C# wrapper is just a collection of P/Invoke functions declared with the
DllImport attribute.
In my first version of this article, I implemented the wrapper functions in a managed C++ assembly, but the resulting DLL was way too large (more than 160 KB). By implementing the wrapper as an unmanaged DLL, with just very little C# glue in a separate assembly, I reduced the combined DLL size to about 35 KB, which is much more reasonable.
Here is the implementation of the wrapper used to get the color of a text element using
uxtheme.dll :
bool __stdcall Wrapper_GetTextColor (const wchar_t* name, const wchar_t* part_name, const wchar_t* state_name, int* r, int* g, int* b) { bool ok = false; if (xpStyle.IsAppThemed ()) { HTHEME theme = xpStyle.OpenThemeData (NULL, name); if (theme != NULL) { try { int part; int state; if (FindVisualStyle (name, part_name, state_name, part, state)) { COLORREF color; int prop = TMT_TEXTCOLOR; if (S_OK == xpStyle.GetThemeColor (theme, part, state, prop, &color)) { *r = GetRValue (color); *g = GetGValue (color); *b = GetBValue (color); ok = true; } } } catch (...) { // Swallow any exceptions - just in case, so we don't crash the caller // if something gets really wrong. } xpStyle.CloseThemeData (theme); } } return ok; }
I derived
System.Windows.Forms.TabPage and wrote the override for method
OnPaintBackground , which called into the wrapper assembly. But there still was a problem I had to address : the text labels I put on the tabbed view did not display the proper background; when setting the control's property
TabPage.BackColor to
Colors.Transparent, the background has not exactly the expected shade. What happens, in fact, when you tell .NET that the background of a control is transparent, is that the erasing code simply calls the parent's
OnPaintBackground method, specifying a clipping region matching the control's shape. When subsequently calling
DrawThemeBackground, the theme gets drawn with the wrong origin (the origin of the clipping region is used, instead of the origin of the containing
TabPage). Code to offset the theme is therefore required...
My first naive attempt was to just offset the background to match the origin of the clipping region. This worked fine as long as the view did not get partially obscured, and then partially repainted : in this case, the clipping origin no longer matched the origin of the label object and the background was, once more, painted with the wrong offset. I finally came up with a rather complex piece of code which walks trough all children, trying to identify which one's background is being erased and using the appropriate offset :
public class TabPage : System.Windows.Forms.TabPage { public TabPage() { } public bool UseTheme { get { return OPaC.uxTheme.Wrapper.IsAppThemed (); } } protected override void OnPaintBackground(System.Windows.Forms.PaintEventArgs e) { if (this.UseTheme) { int ox = (int) e.Graphics.VisibleClipBounds.Left; int oy = (int) e.Graphics.VisibleClipBounds.Top; int dx = (int) e.Graphics.VisibleClipBounds.Width; int dy = (int) e.Graphics.VisibleClipBounds.Height; if ((ox != 0) || (oy != 0) || (dx != this.Width) || (dy != this.Height)) { this.PaintChildrenBackground (e.Graphics, this, new System.Drawing.Rectangle (ox, oy, dx, dy), 0, 0); } else { this.ThemedPaintBackground (e.Graphics, 0, 0, this.Width, this.Height, 0, 0); } } else { base.OnPaintBackground (e); } } private bool PaintChildrenBackground(System.Drawing.Graphics graphics, System.Windows.Forms.Control control, System.Drawing.Rectangle rect, int ofx, int ofy) { foreach (System.Windows.Forms.Control child in control.Controls) { System.Drawing.Rectangle find = new System.Drawing.Rectangle (child.Location, child.Size); if (find.Contains (rect)) { System.Drawing.Rectangle child_rect = rect; child_rect.Offset (- child.Left, - child.Top); if (this.PaintChildrenBackground (graphics, child, child_rect, ofx + child.Left, ofy + child.Top)) { return true; } this.ThemedPaintBackground (graphics, child.Left, child.Top, child.Width, child.Height, ofx, ofy); return true; } } return false; } private void ThemedPaintBackground(System.Drawing.Graphics graphics, int ox, int oy, int dx, int dy, int ofx, int ofy) { System.IntPtr hdc = graphics.GetHdc (); OPaC.uxTheme.Wrapper.DrawBackground ("TAB", "BODY", null, hdc, -ofx, -ofy, this.Width, this.Height, ox, oy, dx, dy); graphics.ReleaseHdc (hdc); } }
I soon discovered that the
System.Windows.Forms.GroupBox did not paint its background with the proper color when using
FlatStyle.System, which is quite annoying. It insisted on painting the background with the default background color. This meant that I had to derive
GroupBox and provide my own
OnPaint implementation, based on
uxtheme.dll. The replacement class I provide can be used both with the classic and the XP styles (and it supports switching from one to the other).
After I started using my own
GroupBox, it became obvious that I had to derive
System.Windows.Forms.CheckBox too, since it too insisted on painting its background itself... Implementing the check box was not really complicated : I just had to make sure that the look and feel was exactly the same as the original version shipped by Microsoft (drawing the caption at the right place, drawing the focus rectangle, reflecting the state of the widget - hot/pressed/disabled/etc.)
If you are interested by the gory details, have a look at the sources...
To use the code provided in this article, you must compile the unmanaged DLL (
OPaC.uxTheme.Win32.dll) and copy the contents of the bin\debug and bin\release directories found in the solution root to your application's bin\debug and bin\release directories. You then add a reference to
OPaC.Themed.Forms.dll (which depends on
OPaC.uxTheme.dll and
OPaC.uxTheme.Win32.dll)...
Create your user interface and then replace
System.Windows.Forms.TabPage with
OPaC.Themed.Forms.TabPage,
System.Windows.Forms.GroupBox with
OPaC.Themed.Forms.GroupBox and
System.Windows.Forms.CheckBox with
OPaC.Themed.Forms.CheckBox. Do not set these object's
BackColor to
Color.Transparent or try to force their
FlatStyle to anything else than
FlatStyle.Standard.
To make sure your controls display as desired when used in a
TabPage, follow these additional rules :
Button.
FlatStyle = FlatStyle.System;
RadioButton.Fl
atStyle = FlatStyle.System;
Label.B
ackColor = Color.Transparent;
Keep the default values for the
FlatStyle and
BackColor properties for all other controls.
I tried out every possible solution I could dream of in order to get the controls to paint properly in a tab control. Only one solution addressed all possible problems, and it is far from elegant. If only Microsoft made the sources of
System.Windows.Forms available (as they did for the Rotor implementation of the .NET framework), fixing the controls would have been simple, efficient and elegant.
A helpful reader hinted me to use function
EnableThemeDialogTexture in
uxtheme.dll, which can be used in Win32 to tell a window how to paint its background, but as expected, it does not work either with the Windows Forms implementation. So the solution presented in this article still seems to be the only way to go.
The source files were last updated on 24 July 2003.
General
News
Question
Answer
Joke
Rant
Admin | http://www.codeproject.com/KB/tabs/themedtabpage.aspx | crawl-002 | refinedweb | 1,491 | 59.6 |
Subject: Re: [boost] [Timer] Inconsistent behaviour
From: Beman Dawes (bdawes_at_[hidden])
Date: 2008-10-29 18:07:10
On Wed, Oct 29, 2008 at 3:15 PM, Bjørn Roald <bjorn_at_[hidden]> wrote:
> Jordans, R. wrote:
>
>>,
>>
>
> If you want a portable wall-clock timer and a simple Boost.Timer like
> interface you can just copy the boost/timer.hpp header file into your own
> namespace and rewrite it to use boost/date_time/posix_time stuff instead of
> ::clock(). I have done that before and it was really simple. There is a
> BOOST_ALL_NO_LIB macro you can define before you include the date_time
> headers so you avoid creating dependencies to anything but the header files
> in boost.
I've got a much better timer with uniform behavior across Windows and POSIX.
But I've been holding off submitting it until the C++0x date-time interfaces
froze. That has now happened, and Howard Hinnant has made his proof of
concept implementation available under the Boost license to use as a
starting point for a Boost implementation. Maybe I can get something good
enough to be useful together sometime in November.
--Beman
Boost list run by bdawes at acm.org, gregod at cs.rpi.edu, cpdaniel at pacbell.net, john at johnmaddock.co.uk | https://lists.boost.org/Archives/boost/2008/10/143999.php | CC-MAIN-2019-26 | refinedweb | 209 | 66.44 |
On 05/02/2008, at 21:39, Adam Atlas wrote: > On 5 Feb 2008, at 18:21, Brett Cannon wrote: >> My current idea is the new names cookie.client and cookie.server for >> Cookie and cookielib, respectively. While this goes against the goal >> of making the new names easier to work with, Cookie has to be renamed >> because of its PEP 8 violation. And having cookie and cookielib in >> the >> stdlib will not help with differentiating between the two. I like this, but maybe it goes against some notion of a shallow stdlib namespace... so maybe cookieclient and cookieserver or something is better. > Has anyone proposed dropping the -lib suffix on *all* modules that > currently use it? It annoys me. It seems arbitrary and redundant. how would you call httplib? httpclient I guess. +50 if I could vote. -- Leonardo Santagada | https://mail.python.org/pipermail/python-ideas/2008-February/001386.html | CC-MAIN-2016-30 | refinedweb | 140 | 77.94 |
#include <db.h>
int memp_register(DB_ENV *env, int ftype, int (*pgin_fcn)(DB_ENV *, db_pgno_t pgno, void *pgaddr, DBT *pgcookie), int (*pgout_fcn)(DB_ENV *, db_pgno_t pgno, void *pgaddr, DBT *pgcookie));
The memp_register function registers page-in and page-out functions for files of type ftype in the specified pool.
If the pgin_fcn function is non-NULL, it is called each time a page is read into the memory pool from a file of type ftype, or a page is created for a file of type ftype (see the DB_MPOOL_CREATE flag for the memp_fget function).
If the pgout_fcn function is non-NULL, it is called each time a page is written to a file of type ftype.
Both the pgin_fcn and pgout_fcn functions are called with a reference to the current environment, the page number, a pointer to the page being read or written, and any argument pgcookie that was specified to the memp_fopen function when the file was opened. The pgin_fcn and pgout_fcn functions should return 0 on success, and an applicable non-zero errno value on failure, in which case the shared memory pool interface routine (and, by extension, any Berkeley DB library function) calling it will also fail, returning that errno value.
The purpose of the memp_register function is to support processing when pages are entered into, or flushed from, the pool. A file type must be specified to make it possible for unrelated threads or processes that are sharing a pool, to evict each other's pages from the pool. During initialization, applications should call memp_register for each type of file requiring input or output processing that will be sharing the underlying pool. (No registry is necessary for the standard Berkeley DB access method types because DB->open registers them separately.)
If a thread or process does not call memp_register for a file type, it is impossible for it to evict pages for any file requiring input or output processing from the pool. For this reason, memp_register should always be called by each application sharing a pool for each type of file included in the pool, regardless of whether or not the application itself uses files of that type.
There are no standard values for ftype, pgin_fcn, pgout_fcn, and pgcookie, except that the ftype value for a file must be a non-zero positive number because negative numbers are reserved for internal use by the Berkeley DB library. For this reason, applications sharing a pool must coordinate their values among themselves.
The memp_register function returns a non-zero error value on failure and 0 on success.
The memp_register function may fail and return a non-zero error for errors specified for other Berkeley DB and C library or system functions. If a catastrophic error has occurred, the memp_register function may fail and return DB_RUNRECOVERY, in which case all subsequent Berkeley DB calls will fail in the same way. | http://pybsddb.sourceforge.net/api_c/memp_register.html | CC-MAIN-2017-47 | refinedweb | 478 | 50.5 |
Byte-oriented applications benefit from an assembly trick
Brenton is the principal software engineer of Grouse Software, based in Adelaide, South Australia. He can be contacted at behoffski@grouse.com.au.
Sidebar: A High-Speed Static Huffman Decoder
Many problems can be efficiently handled with a custom virtual machine. This machine provides an application-specific language that is more powerful, expressive, and concise than any general-purpose programming language. Because it can be carefully tailored to the particular problem domain, a virtual machine can provide a compact, efficient way to solve a broad class of problems. However, simulating virtual machines can be time consuming.
In this article, I'll describe a technique I used for implementing a virtual machine for text processing; this virtual-machine architecture forms the core of both a simple word-count utility and a fast grep program. The grep utility is called Grouse Grep (ggrep). "Grouse" is Aussie slang for "very good" (although most of my acquaintances have noted that the definition of "grouser" is "a grumbler"). Source code and executables for these programs are available electronically from DDJ (see "Availability," page 3) and at.
High-Speed Implementation
Many virtual machines can be implemented as finite-state machines. Each of the machine's states describes how the machine reacts to events while in that state. The fastest way to implement this is to use a table for each state. Each table contains an entry for each event that describes how to handle that event. There is almost no overhead; only a few instructions are required to select the action and dispatch control to the appropriate routine.
You can speed up your implementation further by using threaded code. Instead of storing a code for each action, you just store the address of a handling routine directly in the table. This technique is common in interpreted languages (such as Forth) and is a good method of reducing the cost of handing control to the next action.
On the 80x86 family, you can build compact state tables by using single-byte addresses. Just make sure that all of your handlers start within the same 256-byte code page.
The best way to develop these high-performance machines is to prototype the machine states and actions in a high-level language, keeping an eye on how the assembly version will work, then port the machine to assembly when the machine's operation has been debugged. Many applications require the state tables to be generated dynamically, and it is convenient and efficient to use a high-level language for this process.
Word/Line Count Utility
As an example, I'll show how I converted a prototype word-count utility written in C into a compact, highly optimized 80x86 assembly-language version.
Counting words requires just two states -- a word state and a whitespace state. Each byte is classified as a word character or a whitespace character as part of the table lookup. The machine's actions increment the word or line counters when appropriate. Figure 1 shows the state machine, and lists the actions required.
The C implementation in Listing One is straightforward. To keep the file-system overhead low, the file is brought into memory in fairly large blocks, and the memory is examined directly by the state machine. The machine's state is maintained by a state-table pointer, which points to the whitespace-state table or the word-state table. Word and line counters are also maintained.
The trickiest part is stopping the machine at the end of each buffer. Maintaining a counter virtually doubles the cost of the innermost loop. Instead, I add a marker byte to the end of the buffer, and create additional actions to handle this marker. I've chosen the byte 238 (hex EE), since this value is uncommon in typical text files.
The key part of this state machine is three lines:
NextCh = *pText++; Action = pTab[NextCh]; switch (Action) {
By carefully choosing registers and memory layout, each of these three lines becomes a single assembly-language instruction. Since this sequence is only four bytes, you can simply copy it at the end of each action, avoiding another jump:
lodsb xlat jmp ax
The lodsb/xlat sequence uses si (pText) and bx (pTab) to load the corresponding action into the lower byte of ax. As long as all of the actions are in the same 256-byte page, you can initialize the upper part of ax outside the inner loop.
The assembly-language version of the word-count machine is also available electronically. Note that the values chosen to represent the actions in the C program are the entry points of the corresponding assembly-language action. The word count runs twice as quickly as the C version, but the utility is heavily limited by the file-system interface, spending more time waiting for the file to be read than it does actually performing the count.
Regular Expressions
A more sophisticated example of this approach is Grouse Grep, my implementation of the grep search utility. By compiling the regular expression (RE) into a fast state machine, my program runs significantly faster than traditional recursive search implementations.
I'll start by showing how to compile a regular expression into a simple, unoptimized state machine, and then discuss some optimizations. The basic regular expression match is performed on a line terminated with NUL.
Matching Individual Bytes
The simplest search case is finding a single byte (such as "x") in the line. This search requires a one-state machine with three actions -- COMPLETED, ABANDON, and AGAIN. The state describes how the machine operates while it is scanning for the desired byte.
The table entry for "x" contains the action COMPLETED, which stops the search and reports the current position as the successful matching text. The table entry for NUL contains the action ABANDON. Since NUL marks the end of the line, this action will stop the search if no match is found.
All other entries in the table contain AGAIN, which instructs the search engine to move to the next character of the line, translate the byte into the action in the state table, and perform the specified action. This action makes the search engine scan the line from left to right.
Example 1 describes the state table for this search, and provides both C and Intel 80x86 assembly versions of the search engine.
Matching Classes
Matching classes of characters merely involves writing COMPLETED for each character of the table that matches. For example, a state table for matching the class "[aeiou]" is implemented by writing COMPLETED in the "a", "e", "i", "o", and "u" table entries, along with ABANDON for NUL and AGAIN for the remaining entries. The same idea applies to handling case insensitivity and the period character (which matches anything).
Timing information for the 486 shows that the cost of the lodsb/xlat/jmp sequence is nominally 14 clock cycles. This is slightly more costly than conventional code when matching a single byte, but is cheaper than any alternative code for matching classes of bytes.
Backtracking
The examples so far contain only a single search state. Searching for a string such as "koala" requires a search engine with at least two states: a "search for k" state and a "match oala" state. If a "k" is found, the search advances to the next state and looks for "oala".
If the match attempt fails, the search must backtrack to the initial state and continue searching for the next "k". The search engine must remember the text position and the match state for each point where the search may resume after failing to complete a match attempt.
Most existing regular expression searches use recursion to maintain backtracking information. While simple and effective, this is slower than the table-driven machine, because of the bookkeeping associated with function calls and returns. Most regular expression programs try to avoid recursion where possible because of this expense. The finite-state machine implementation eliminates this overhead entirely.
When a backtracking pathway is found that may be required later in the search, a text/state table pointer pair (describing how to resume) is pushed onto a stack. Backtracking from any state is then a simple matter of popping the pair off the stack and invoking the next threaded action according to the state table.
An additional bonus for the assembly-language version is that the CPU stack can be used directly to maintain the backtracking, allowing the stack to be implemented very efficiently.
String Matches
Rather than implement a "match string" action, it's more convenient to give each character of the search string its own state, to allow multicharacter match cases in any position. Three additional search engine actions are required:
- ADVANCE, which advances to the next state for the next text character when the current state matches.
- START_PUSH, which pushes a backtracking state/text context when "k" is found so that the "k" search can resume if the "oala" scan is unsuccessful.
- NO_MATCH, which reverts to the last backtracking state/text context, executed during any of the "o", "a", "l", or "a" states if no match found.
Figure 2 shows the state machine for this search, including arrows for the various actions possible in each state. It also introduces an ok and fail state. These states handle the ultimate success or failure of the search, and allow the remaining states to perform matches without any line management overhead. Every entry of the fail table contains ABANDON; every entry of the ok table contains COMPLETED. The backtracking stack is initialized with a fail reference to catch the underflow when all backtracking options are exhausted.
Anchor to Start or End
To anchor the search to the start of the line, merely change the AGAIN actions in the first match state to ABANDON. If the first state does not match the first character of the line, the match fails, without the engine searching the remainder of the line for a possible match start.
Anchoring the search to the end of the line is easy to implement by editing the ok state so that almost all bytes contain NO_MATCH, and only end-of-line characters such as LF, CR, and NUL contain COMPLETED.
Optional Elements
Optional elements are implemented by adding two actions to the engine. These actions convert dead ends into match paths, and add an optional path to the backtracking stack if the match succeeds. In both cases, the option is included if available, but the match can advance if it is not. See Figure 3, which shows the state tables to search for "ding?o".
Where the input matches the optional state (e.g., "dingo"), an additional match path omitting the state is pushed as the match advances. The path pushed consists of the next state plus the previous character. Where the first path attempted fails to match, the alternative is selected by the backtracking mechanism.
If the optional element isn't matched (e.g., "dino"), the BACK_AND_ADVANCE action instructs the machine to retry the match with the next state.
The undo behavior is achieved by adding a "zero trip" case to the backtracking stack as part of the match. The next state and the previous character are added to the backtracking stack, so that if the first match path fails, the alternative path, which effectively undoes the optional match, is attempted. Similarly, when matching "quoka" using "quokk?a", the "a" does not match the second "k", but because the action is BACK_AND_ADVANCE instead of NO_MATCH, the "a" is retried in the next state, leading to a successful match.
Iteration
Iteration is specified by "*", meaning "zero or more," and "+", meaning "one or more." "*" iteration repeats the previous RE zero or more times: "xy*z" matches "xz", "xyz", "xyyz", "xyyyz", and so on. The "+" operator is the same, except that the previous expression is duplicated before the iteration, so that "a[bc]+d" is equivalent to "a[bc][bc]*d". Once this substitution is made, both cases can be treated identically. The correct method of handling iteration is to try to move forward as far as possible within the iteration state, and push option points into the backtracking stack as the loop proceeds. If the iteration fails, revert to the last item on the backtracking stack and try to match the states after the iteration again.
Figure 4 shows the additional action required for iteration, and displays the tables and actions to search for "bil*by".
For the text "biby", the "l" match state fails, but the BACK_AND_ADVANCE action means the "b" is retried in the fourth state, leading to a successful match. For the text "billlllby", the "l"s are matched and the match stays in the same state. Again, the "b" does not match the iteration but is matched in the next state.
This completes the minimal "language" required to perform basic RE searches. This version usually runs faster than recursive implementations, but slower than fixed-string searches.
The code to map the RE from the plain text specification to the state tables was written in C. The search engine was prototyped in C and then ported to assembly. The assembly source to implement this language is available electronically.
Optimizing for Speed
The speed of grep is not the speed of the regular expression match -- it is the speed at which the search engine can discard lines that can't match the RE. Improving the speed of the search breaks down into the following areas:
- Minimizing the effort required to read the file from the operating system and break the data into lines.
- Optimizing searching by using string-scanning techniques such as the Boyer-Moore optimization or by exploiting the CPU's search instructions.
- Editing the search order to look for easy parts of the RE first so that lines that can't match are eliminated quickly.
- Analyzing relationships between tables and optimizing state transitions (especially iterative backtracking).
File Buffering
The main way to reduce both the file-system cost and the line-splitting cost is to read the file directly into a large buffer in memory, and to adapt the search algorithm to recognize and handle line separators as they are encoded in the file. However, the file is often larger than the memory buffer. The solution is to read as much of the file as possible, then trim the buffer so it only contains complete lines. The file is rewound to the trim point, so that the next buffer will start at the start of a line.
LF is treated as the line separator for counting and display purposes, and CR also acts as a line terminator for searches anchored to the end of the line. Since the NUL is no longer needed as a line terminator, it is used as a buffer terminator. This allows almost all search actions to ignore the end-of-buffer cases.
The search optimizations are implemented by inserting an additional machine state between the fail state and the first expression state, called the "start state." This state receives actions specific to the speed optimizations without disrupting the actions of any of the existing expression search states.
String Searching
Some options of grep, such as selecting lines that don't match, work much better when the file is presented as a series of lines. Other options, such as line-number reporting, require that the start search examine each byte. However, simple searches can afford to skip bytes during the match start search, allowing the use of the Boyer-Moore algorithm (see "A Fast String Searching Algorithm," by R.S. Boyer and J.S. Moore, Communications of the ACM, October 1977). This technique is the main method used by other searches to gain speed.
The Boyer-Moore algorithm is simple: If you are searching for a simple string such as "galah", don't look in the first position for "g" -- look in the fifth position for "h". If you find an "h", start a normal match attempt from the first character. If you don't find an "h", look at the character you inspected: If it isn't in the string at all, skip forward by the length of the string. Even when the character is in the string, the search can still skip forward: For example, finding an "l" when searching for "galah" allows the search to skip forward two bytes.
A series of skip-byte actions were added to the search engine. The optimized search for "galah" has SKIP4 for "g", SKIP2 for "l", SKIP1 for "a", and START_OFFSET_MATCH for "h", with all other characters having SKIP5, except for the zero byte, which has ABANDON. In general, the skip size for an element is found by counting the number of NO_MATCH entries for that element in the tables leading up to the last state of the string.
CPU Byte Search
Scanning for a single byte is best done by using the processor's byte-search instruction. On the 486, this instruction scans memory twice as quickly as the lodsb/xlat/jmp sequence. A BYTE_ SEARCH action is used in the starting state where the Boyer-Moore algorithm is not available because the string is only a single literal ("%", for example) or because the search contains elements that cannot be optimized after the initial literal ("x.*y.*z").
Easiest First
Some regular expressions, such as "x[^a]*[a-z]*[^c]*wombat" are very inefficient if searched from left to right, as the first elements require massive amounts of backtracking. However, these expressions usually have an easier component that can be searched efficiently -- in this case, "wombat". Ggrep looks for easy strings in the RE, and splits the RE in two if a worthwhile position is found. The optimization is implemented by analyzing and editing the regular expression before attempting to produce the table-driven version.
In the example, the code searches for "wombat", and, where found, splits the line and searches for "x[^a]*[^b]*[^c]$" (anchored to the end of the line to ensure the split search may be reassembled). The line matches if both parts of the RE match, and the text selected by each match is combined to produce the final match.
Some care is required when implementing this optimization to ensure that the search continues examining the current line if the first part of the regular expression does not match. In addition, extra effort is required if the match is to report the longest possible part of the line matching the specified expression.
Optimizing Iteration
The amount of backtracking generated by iteration states can be significantly reduced by inspecting the next state after the iteration. Consider the expression "rock.*wallaby". While looping through the ".*" state (which matches everything up to the end of the line), it is easy to see that proceeding to the next state is only worthwhile if the character is "w": Any other character will not match.
These redundant backtracking cases can be eliminated by replacing AGAIN_ PUSH_ADVANCE with AGAIN for every element of the iteration table where the next table has NO_MATCH. This optimization is done as a post-processing step once the expansion of the search into tables is complete.
The result is that the search does not incur the cost of pushing the redundant backtracking paths, and (more importantly) reduces the number of backtracking paths that may need to be considered. In many cases, this optimization raises the performance of iterative searches close to that of simple string searches.
Acknowledgments
Thanks to Ross Williams, David Knight, Christian Adami, Ted Bullen, Martin Hogan, Mark Rawolle, and Simon Hackett for reviewing a draft of this article and providing valuable suggestions. Thanks also to my family for their support.
Listing One
#include "fcntl.h"#include "io.h" #include <stdio.h> </p> typedef unsigned short UINT16; /*Unsigned 16-bit integer*/ typedef unsigned char BYTE; /*unsigned byte*/ </p> #define PAGE_ALIGNED far #define ENDMARKER 0xEE </p> #define EE_NOP 0x00 /*Buf end check, NOP if not*/ #define NOP 0x04 /*Stay in same state*/ #define EE_WORD 0x08 /*Buf end check, WORD if not*/ #define WORD 0x0c /*Words++, change to word*/ #define WHITE 0x14 /*Change to whitespace*/ #define LF_WHITE 0x1a /*Lines++, change to whitespace*/ #define LF_NOP 0x1c /*Lines++, stay in same state*/ #define BUF_SIZE (28*2048u) </p> /*Global variables updated directly by counting loop*/ unsigned long Lines; unsigned long Words; BYTE *pCurrentTable; </p> BYTE Tables[512]; #define WhiteTable (&Tables[0]) #define WordTable (&Tables[256]) </p> static BYTE Buf[BUF_SIZE + 2]; static void InitTables(void) { int i; for (i = 0; i < 256 ; i++) WhiteTable[i] = WORD; WhiteTable[ 9] = WhiteTable[11] = WhiteTable[12] = NOP; WhiteTable[13] = WhiteTable[32] = NOP; WhiteTable[10] = LF_NOP; WhiteTable[ENDMARKER] = EE_WORD; </p> for (i = 0; i < 256; i++) WordTable[i] = NOP; WordTable[ 9] = WordTable[11] = WordTable[12] = WHITE; WordTable[13] = WordTable[32] = WHITE; WordTable[10] = LF_WHITE; WordTable[ENDMARKER] = EE_NOP; } /*C and assembly versions of word counter*/ void PAGE_ALIGNED CountC(UINT16 NrBytes, BYTE *pText); void PAGE_ALIGNED CountAsm(UINT16 NrBytes, BYTE *pText); static int WCFile(char *pFilename); int main(int argc, char **argv) { InitTables(); while (--argc) { /*Count each file*/ if (! WCFile(*++argv)) return 1; } return 0; } static int WCFile(char *pFilename) { int NrBytes; int Handle; /*Attempt to open file for reading as raw bytes*/ Handle = open(pFilename, O_BINARY | O_RDONLY); if (Handle == -1) { printf("\nCan't open: %s", pFilename); return 0; } Lines = 0; Words = 0; pCurrentTable = WhiteTable; do { /*Read a slab of the file into memory*/ NrBytes = read(Handle, Buf, BUF_SIZE); /*Exit loop if there's an error*/ if (NrBytes == -1) break; /*Process the slab using CountC or CountAsm*/ CountAsm(NrBytes, Buf); /*Stop if finished with file*/ } while (NrBytes >= BUF_SIZE); printf("%7lu %7lu %7lu %s\n", Lines, Words, tell(Handle), pFilename); (void) close(Handle); return 1; } /*WCFile*/ void PAGE_ALIGNED CountC(UINT16 NrBytes, BYTE *pText) { BYTE NextCh; BYTE Action; BYTE *pTab = pCurrentTable; BYTE *pEnd = &pText[NrBytes]; /*Add endmarker to buffer to trigger end test below*/ *pEnd++ = ENDMARKER; /*Loop through each byte*/ for (;;) { NextCh = *pText++; Action = pTab[NextCh]; switch (Action) { case EE_NOP: if (pText == pEnd) break; /*FALLTHROUGH*/ case NOP: continue; </p> case EE_WORD: if (pText == pEnd) break; /*FALLTHROUGH*/ case WORD: Words++; pTab = WordTable; continue; case WHITE: pTab = WhiteTable; continue; case LF_WHITE: pTab = WhiteTable; /*FALLTHROUGH*/ case LF_NOP: Lines++; continue; } /*End of buffer found: remember in-word context and exit*/ pCurrentTable = pTab; return; } } /*CountC*/ </p> </p>
DDJ | http://www.drdobbs.com/parallel/high-speed-finite-state-machines/184410318 | CC-MAIN-2014-52 | refinedweb | 3,721 | 57.81 |
Python program to extract a single value from JSON response (Using API call)
Hello Everyone! In this Python tutorial, we are going to learn how to retrieve Single data or single values from JSON using Python. To perform this task we will be going to use the request module in Python, this module allows users to send HTTP requests and receive responses in form of JSON.
How to extract a single value from JSON response
Let’s start by importing the requests module,
import request import urllib.parse
After importing modules,
import urllib.parse import requests base_url=" your API key here/pair/" print("Enter the First Currency") s=input() print("Enter the Second Currency") l=input() value=s+"/"+l url = base_url+value json_data = requests.get(final_url).json() result = json_data['conversion_rate'] print("Conversion rate from "+s+" to "+l+" = ",result)
- Declare base_url with API key variable.
- Take the inputs from the user.
- Add the user input to our base_url and make final_url or make an API request to the server and fetch the data from the server.
- Now, json_data makes an API call and fetches the data from the server & it contains JSON response.
- We will get the result from the website in JSON format.
- So let us create a variable called result which will contain the JSON data and retrieve the single data which is required.
- To retrieve single data like ‘conversion_rate’ you have to declare a variable from the JSON response.
- The ‘result’ variable holds the value of ‘conversion_rate’.
- Final print the result.
JSON RESPONSE
result "success" documentation "" terms_of_use "" time_last_update_unix 1615075202 time_last_update_utc "Sun, 07 Mar 2021 00:00:02 +0000" time_next_update_unix 1615161617 time_next_update_utc "Mon, 08 Mar 2021 00:00:17 +0000" base_code "USD" target_code "INR" conversion_rate 73.0648
OUTPUT
Enter the First Currency USD Enter the Second Currency INR Conversion rate from USD to INR = 73.0648
Now, you can understand how to retrieve single data from a variety of other APIs. | https://www.codespeedy.com/python-program-to-extract-a-single-value-from-json-response-using-api-call/ | CC-MAIN-2021-17 | refinedweb | 322 | 54.52 |
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I need help. "The maximum value of day depends on the month for the date. The maximum day for the month of February also depends on the year value for the ExtendedDate object (i.e. is it a leap year...
Type: Posts; User: Marcusjamaal
I need help. "The maximum value of day depends on the month for the date. The maximum day for the month of February also depends on the year value for the ExtendedDate object (i.e. is it a leap year...
Here's the swing class.
package murach.forms;
import javax.swing.*;
public class SwingValidator
need help with this GUI. Can't get the calculate button to actually calculate. When I press it nothing happens. Only button that works is the exit button.
import javax.swing.JFrame;
I'm sorry jps, I'm a little confused, do you mean where's the loop that will print each entry each time it's entered?
I was under the impression from the instruction I was given for the assignment...
Hi, having trouble with this code. Can't figure out how to tell the code to keep printing firstname lastname and etc.. based on what number of students the user enters and how to keep storing it.
...
nvm guys, I apologize. i messed with it enough and figured it out. i couldn't use a package with java as the name.
package newstuff;
import java.text.NumberFormat;
import java.util.Scanner;
...
Now it's giving me errors I don't understand.
java.lang.SecurityException: Prohibited package name: java
at java.lang.ClassLoader.preDefineClass(ClassLoader.java:649)
at...
I have a question about my code I'm trying for the exercise.
It says "Error, cannot load ValidatedInvoiceApp." What's wrong?
package java;
import java.text.NumberFormat;
import...
Thanks Z! Wow, this is gonna help me a ton!
I'm having these problems now with it.
1. theres a problem with "else without if"
2. while ( i >= 1) , cannot find symbol <identifier expected>
3. Alot of number 2 for mutiple functions
Z
heres an updated code...I can say I'm understanding you, but it's alot different then actually doing so...
public static void main(String[] args) {
// welcome the user to...
Thank you Z! :o. Also yes I do. I'm confused by your message though, if I'm prompting the user to enter input for the value of the numbers, my code must differ from your code to get the first and...
Sorry Guys here's what's specifically wrong with it
1. Illegal start of expression , cannot find symbol, symbol class : class first number, location: class Greatest common division finder.
...
Thank you. :o
Hi everybody I'm extremely new to java programming and was asked to do this >
" The formula for finding the greatest common divisor of two positive integers x and y
follows the Euclidean... | http://www.javaprogrammingforums.com/search.php?s=1391c48d724a9df0c7300151a44f0294&searchid=837318 | CC-MAIN-2014-15 | refinedweb | 477 | 69.38 |
I...
I've seen many Win32 Registry API wrapper classes
but all of them failed to actually make it easier to use the registry than
the API. Most have been simple wrappers which add no value (apart from saving
you the bother or having to pass the HKEY to each function...).();
}
RegDeleteKey behaves differently on Win95 and NT. CRegistryKey::DeleteKey( )
acts consistently on both platforms, never removing sub keys, whilst CRegistryKey::DeleteKeyAndSubkeys( )
always removes sub keys.
RegDeleteKey
CRegistryKey::DeleteKey(
CRegistryKey::DeleteKeyAndSubkeys(
In addition, CRegistryKey::OpenKey( ) only opens a key and does not create
one, use CRegistryKey::CreateKey( ) to create one and CRegistryKey::CreateOrOpenKey( )
if you don't care.
CRegistryKey::OpenKey(
CRegistryKey::CreateKey(
CRegistryKey::CreateOrOpenKey(
CRegistryKey is slightly more complex than just a wrapper. It's actually
implemented using the "handle-body idiom", the CRegistryKey
object is a handle only the underlying HKEY which is reference counted. This
is so that it's easy to pass CRegistryKey 's by value. Since CRegistryKey
owns an open HKEY, and since that key is closed when the CRegistryKey
object goes out of scope we have to be careful when copying them around. It's
not enough to allow the default copy constructor and assignment operators
to be used. Allowing that would mean that the underlying HKEY could be closed
more than once, as in the example below...
void DoStuffWithKey(CRegistryKey key1)
{
// blah blah blah, do clever stuff with a key...
// key1 is closed here when it goes out of scope!
}
try
{
CRegistryKey key1 = CRegistryKey(HKEY_LOCAL_MACHINE, _T("SOFTWARE"));
// call a function and pass the key by value...
DoStuffWithKey(key1);
// key1 is closed here - for the second time!!!
}
catch (CRegistryKey::Exception &e)
{
e.MessageBox();
}
An alternative could be to duplicate the HKEY using DuplicateHandle( )
in the copy constructor and assignment operator, but that's messy, and unnecessary,
and besides it will fail if the HKEY is a key on a remote machine.
A better, if slightly more complex, solution is to hold a single representation
of the underlying HKEY and only close it when the last CRegistryKey
that references it is destroyed.
HKEY
DuplicateHandle(
To achieve this, we store a counter with the HKEY and increment the
counter every time we copy a CRegistryKey that refers to the HKEY ,
and decrement the counter when we destroy such a CRegistryKey . With
this in place there is only ever one copy of the open HKEY no matter
how many additional CRegistryKey objects are created from the initial
one. What's more, as a user of CRegistryKey we never see any of this
complexity - it just works.
The code that makes it easy to navigate sub keys or values associated with
a CRegistryKey is structured in a way that is similar to STL iterators.
This makes it easy and intuitive to use, just ++ your way along the list of
keys or values. When I was writing these iterators, I realised how much code
was similar between the sub key iterator and the value iterator. At first
I tried to factor this common code into a common base class. This didn't really
work, a lot of the common stuff was in assignment or equality operations which
didn't work well as base class members. Next I tried a template base class
which took the derived class with the specialist knowledge as its template
parameter. Whilst this was better (the assignment and equality ops could be
based on the template parameter rather than the base class) it still wasn't
ideal as constructor of the iterator has to "prime the pump" by
advancing to the first available item. This involved calling the Advance()
function which in turn needed derived class functionality in the form of a
virtual function call to GetItem() . Of course, a virtual function call
from a base class constructor is not going to work in this instance.
Advance()
GetItem()
The end result was an iterator template which
takes a template parameter to its base class. The base class implements the
specialist knowledge and is available at any point in the life of the template
derived class. To make life easier for myself I write a very simple abstract
interface class which represented the interface required by the template class
for it to work. This wasn't required, the template would have worked with
any class that implemented the required interface, not just those derived
from the abstract interface class, but it made the requirements easier to
see. Base classes are quite at liberty to add any additional functionality
that they require of the resulting iterator - for example, they might add
access functions for parts of the item being iterated over. While I was implementing
the iterator requirements base class I realised that all of the handling of
the underlying registry key, which was common between both implementations,
would need to go here, rather than in the template class. Only if the code
were in a base class of the implementation could the implementation use it...
The resulting classes involved look something
like this...
At the end of the day, for the sake of removing some duplicate code, the iterators
are more complex. This complexity is purely an implementation issue, it doesn't
leak through into the iterator interface and affect users. Even so, I'd probably
do them differently next time...
Now that we have examined all of the pieces our
simple Win32 Registry API wrapper class has turned out to be a collaboration
of quite a few classes, as can be seen from the diagram below:
Only CRegistryKey itself is declared at the namespace level, all other
classes are nested within CRegistryKey . This allows for class names
to be less specific as they are already scoped within CRegistryKey
- for example we can safely call the value class "Value" as the
only way to access it is as " CRegistryKey::Value ".
CRegistryKey::Value
At present CRegistryKey doesn't provide wrappers for the following
Win32 Registry API calls, this shouldn't cause a problem since you can always
access the underlying HKEY to make these calls.
RegQueryMultipleValues(
RegQueryValueEx(
RegSetValueEx(
REG_LINK
REG_MULTI_SZ
REG_RESOURCE_LIST
RegQueryInfoKey(
Since a CRegistryKey represents an open registry key there's no way
to close the key whilst the key is in scope. This hasn't proved to be a problem
for me, but if it does you can always assign one of the standard key handle
values to your CRegistryKey which will cause the open key to be closed
and the standard key to be open - since the standard key handle values appear
to be treated differently to normal keys (you don't need to open or close
them) this has the desired effect.
try
{
CRegistryKey key = CRegistryKey(HKEY_LOCAL_MACHINE, _T("SOFTWARE"));
// Do some key stuff...
// Force the key closed...
key = HKEY_LOCAL_MACHINE;
// Do other stuff in this scope...
}
catch (CRegistryKey::Exception &e)
{
e.MessageBox();
}
I personally tend to scope the key so that it remains open whilst it is in
scope and then is automatically closed when it goes out of scope... Don't
pass the CRegistryKey to the RegCloseKey( ) API as this will
cause the key to be closed twice, once by the RegCloseKey( ) call and
once when the CRegistryKey goes out of scope or has another key assigned
to it.
RegCloseKey(
See the article on Len's homepage for the latest. | https://www.codeproject.com/articles/345/registry-api-wrapper?fid=383&df=90&mpp=25&sort=position&spc=relaxed&tid=2057369 | CC-MAIN-2017-09 | refinedweb | 1,216 | 57.5 |
Q: Why does your book Android Programming heavily advocate the use of fragments when developing Android applications? You can develop Android applications without using any fragments, so why bother? Why do fragments matter?
A: As Android developers, we have two main controller classes available to us: Activities and Fragments.
Activities have been around for a very long time and are used to construct a single screen of your application. When someone is using your Android application, the views that the user sees and interacts with are hosted and controlled by a single activity. Much of your code lives in these Activity classes, and there will typically be one activity visible on the screen at a time.
As an example, a simple activity may look like the following:
public class MainActivity extends Activity { @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_main); } }
When Honeycomb was released, the Fragment class was introduced. Fragments are another type of controller class that allows us to separate components of our applications into reusable pieces. Fragments must be hosted by an activity and an activity can host one or more fragments at a time.
A simple fragment looks similar to an activity, but has different life cycle callback methods. The fragment below accomplishes the same thing as the activity example above: it sets up a view.
public class MainFragment extends Fragment { @Override public View onCreateView(LayoutInflater inflater, ViewGroup container, Bundle savedInstanceState) { return inflater.inflate(R.layout.fragment_main, container, false); } }
As you can see, activities and fragments look very similar and are both used to construct applications. So, what are the advantages that come with the use of Fragments?
Device Optimization
As I mentioned before, fragments were introduced with the Honeycomb release of Android. Honeycomb was the first version of Android with official support for tablets. One of the original goals of the fragment concept was to help developers make applications that provide different user experiences on phones and tablets.
The canonical example of fragment use is an email client. Let’s take a look at the Gmail app. When running on a phone, the Gmail app will display a list of the user’s emails. Tapping on one of these emails will transition the screen to showing the detail view for that particular email. This process is composed of two separate screens.
When using the same app on a tablet, the user will see both the email list and the detail view for a single email on the same screen. Obviously, we can show much more information at once with a tablet.
Fragments make implementing this functionality easy. The list of emails would exist in one fragment, and the detail view for an email would exist in another fragment. Depending on the device size, our activity will decide to show either one or both of these fragments.
The beauty of fragments here is that we do not have to modify either of these Fragment classes for this to happen. Fragments are reusable components of our application that can be presented to the user in any number of ways.
So, this is great, but what if you aren’t developing an app that works on phones and tablets? Should you still use fragments?
The answer is yes. Using fragments now will set you up nicely if you change your mind in the future. There is no significant reason not to start with fragments from the beginning, but refactoring existing code to use fragments can be error-prone and will take time. Fragments can also provide other nice features besides UI composition.
ViewPager
For example, implementing a ViewPager is very easy if you use fragments. If you’re not familiar with ViewPager, it’s a component that allows the user to swipe between views in your application. This component is used in many apps, including the Google Play Store app and the Gmail app. Swiping just swaps out the fragment that the user is seeing.
ViewPager hits at one of the core ideas behind fragments. Once your application-specific code lives in fragments, those fragments can be hosted in a variety of ways. The Fragment is not concerned with how it’s hosted; it just knows how to display a particular piece of your application. Our activity class that hosts fragments can display one of them, multiple Fragments, or use some component like a ViewPager that allows the user to modify how the fragment is displayed. The core pieces of code in your app do not have to change in order for you to support all of these different display types.
Other Features
So, Fragments can help us optimize our applications for different screen sizes and they help us provide interesting user experiences. There are other features that Fragments can help with as well.
You can think of Fragments as a better-designed version of an activity. The life cycle of a Fragment allows for something that Activities do not: Separation of the creation of the View component from the creation of the controller component itself.
This separation allows for the existence of retained fragments. As you know, on Android, rotating your device is considered a configuration change (there are a few other types of configuration changes as well). This configuration change triggers a reload of Activities and Fragments so that resources can be reloaded for that particular configuration. During the reload, these instances are completely destroyed and then recreated automatically. Android developers have to be conscious of this destruction because any instance variables that exist in your activity or fragment will be destroyed as well.
Retained fragments are not destroyed across rotation. The view component of the fragments will still be destroyed, but all other state remains. So, fragments allow for the reloading of resources across rotation without forcing the complete destruction of the instance. Of course, there are certain situations where this is and is not appropriate, but having the option to retain a fragment can be very useful. As an example, if you downloaded a large piece of data from a web server or fetched data from a database, a retained fragment will allow you to persist that data across rotation.
Additional Code
When you use fragments to construct your Android application, your project will contain more Java classes and slightly more code than if you did not use fragments at all. Some of this additional code will be used to add a fragment to an activity, which can be tricky if you don’t know how the system works.
Some Android developers, especially those who have been developing Android applications since before Honeycomb was released, opt to forego using Fragments. These developers will cite the added complexity when using Fragments and the fact that you have to write more code. As I mentioned before, the slight overhead that comes with fragment use is worth the effort.
Embracing Fragments
Fragments are a game changer for Android developers. There are a number of significant advantages that come with Fragment-based architectures of Android applications, including optimized experiences based on device size and well-structured, reusable code. Not all developers have embraced fragments, but now is the time. I highly recommend that Android developers make generous use of fragments in their applications. | http://www.informit.com/articles/article.aspx?p=2126865 | CC-MAIN-2018-17 | refinedweb | 1,209 | 54.02 |
Asked by:
Avoid Registering DLL using GACutil
Working on .Net 3.0.
We have implemented the script task under SSIS to download and unzip the file and which is running fine on our local environment when we tested.
On server environment we are getting the issue “Could not load file or assembly 'Interop.Shell32, Version=1.0.0.0, Culture=neutral, PublicKeyToken=aad7673e5ba23c29' or one of its dependencies.“
This issue occurs when SSIS is unable to load the “Interop.Shell32.dll” file, we can resolve this issue by registering the .dll in GAC in my local environment as mentioned in below steps.
As we don’t have machine access of Server Environment, we are unable to perform those steps.
Is there any other way to dynamically load the dll?
Steps to Register DLL
Go to Project->st_xxxxx Properties, Signing, tick sign the assembly, select New, type any file name, untick Protect my file with a password.
Go to References, Add, COM, select Microsoft Shell Controls And Automation, OK.
** You should see a popup saying No template information found..., just ignore that. **
Highlight the Microsoft Shell Controls you just added and look at the bottom right properties section, make sure "Strong Name" is True.
(if you haven't done step 2., this should be False, redo step 2 and make sure this is True)
Open My Computer and go to the path of the generated Microsoft Shell Controls dll (see the highlighted Path, usually it is C:\Users\xxxxxxx\AppData\Local\Temp\SSIS\yyyyyyyyyyyyyyyyy\obj\Debug\Interop.Shell32.dll). Copy it to somewhere else (its because the original path yyyyyyyyyyyyyy is randomly generated everytime you edit your script, causing that the signed dll gone everytime).
Back to the Project->st_xxxxx Properties, remove the highlighted Microsoft Shell Controls. Click Add, Browse, select the Interop.Shell32.dll you just copied, OK. Again, make sure the Strong Name property is True. If not, redo from step 2.
Still at References, under Imported namespaces, scroll down to tick the box next to Shell32.
Go to Signing, untick Sign the assembly. Go to Project Explorer on top right, delete the key file (xxxx.snk) you created at step 2 (delete the key file within Project Explorer, not from your folder, make sure it is not included in your Project Explorer anymore or your script task wont run).
Save all and close the Script Editing window.
Locate "gacutil.exe" in your computer. You should have it together with the installation of SSIS. Its under C:\Program Files (x86)\Microsoft SDKs\Windows\v7.0A\bin\x64. Type gacutil -i "path of your copied Interop.Shell32.dll in step 5".
Question
All replies
Do you have administrator privileges on that server? Because that is required for gacutil...
Please mark the post as answered if it answers your question | My SSIS Blog: | Twitter
No I dont have
Then either ask the administrator to install the dll in the gac or you have to use other solutions that don't require dll's.
For downloading:
For unzipping you could use av Execute Process Task if there is a zip utility available on your server.
Please mark the post as answered if it answers your question | My SSIS Blog: | Twitter
In .net 3.0 is there any other solution in SSIS to unzip the file without using any third party tool
ZipArchive class is only available in 4.5
In 3.0 you only have Gzip And therefore people often use third party libraries like DotNetZip and sharpziplib. I used DotNetZip in my SSIS unzip task
Please mark the post as answered if it answers your question | My SSIS Blog: | Twitter
- Edited by SSISJoostModerator Friday, November 01, 2013 10:16 AM
- Proposed as answer by Mike YinMicrosoft contingent staff, Moderator Tuesday, November 05, 2013 4:38 PM | http://social.technet.microsoft.com/Forums/en-US/d1263ff0-5f6f-4a77-9e5f-a7183948aef3/avoid-registering-dll-using-gacutil?forum=sqlintegrationservices | CC-MAIN-2014-10 | refinedweb | 635 | 65.32 |
How secure is Docker?
Discussion about Docker security and docker vulnerabilities.
Over the past few years, the use of virtualization technologies has increased exponentially. Accordingly, the demand for efficient and secure virtualization solutions is increasing. The two most used virtualization technologies are Container-based virtualization and hypervisor-based virtualization. Of these two types, container-based virtualization is more lightweight as compared to hypervisor-based virtualization and it provides an efficient virtual environment, but containerization can lead to few security concerns.
So in this article, we will look into different types of security concerns and we’ll analyze the security level of Docker. Docker is a well-known representative of container-based approaches and the scope of this article is limited to security threats in docker containers. In this article, let’s talk about docker security and its vulnerabilities but first, understand what is the docker? and for that, I want to introduce docker to all of you.
Docker introduction :
Docker is an OS virtualization software platform that is used to create, deploy and run applications in a Docker container. It is very lightweight when compared to a virtual machine and it allows us to wrap up our application in a container along with all the required libraries and other dependencies and deploy it, so it reduces dependency-related conflicts. Docker helps in the simplification and acceleration of workflow.
Basically, Docker is nothing but an environment where we can run our code or program and it will run in isolation. Generally, when we do some project in docker it isolates which is not the case in a personal machine. For example, if I am doing the project in system-A and need to show the project in system-B then there are chances that our program might get crashed because of various reasons like different compilers, different versions of the software/packages, or different operating system. So docker helps us to keep the project and all the required libraries in a single package and if we need to export our project to a different system then docker will export the entire environment of our project.
Empowering App Development for Developers | Docker
What's New DockerCon 2021 Call for Presentations now open Share your insight, experiences and best practices with…
Two components of docker:
- Docker Image.
- Docker Container.
Docker defines the Images as follows:
-.
- Image is an executable package that includes everything needed to run an application -the code, runtime, libraries, environment variables, and configuration files.
Docker defines the containers as follows :
A container is a standard unit of software that packages up code and all its dependencies. So, the application runs quickly and reliably from one computing environment to another. The container is an environment which has all package and all the system required for running the program.
Most people get confused in a docker container and virtual machine. These both are very different. In the virtual machine, we need to install operating system + kernel + drivers but docker uses the host’s system, kernel and driver so it's very lightweight and takes very little space. Now-a-days, docker containers are so much in demand because they are flexible, lightweight, interchangeable, portable, scalable, stackable.
How docker isolates itself?
Linux Cgroups is one of the kernel features provided by Linux, which helps docker to isolate itself. Cgroups allow us to group a set of related processes which can run as a single unit. We can control this group of processes in terms of usage CPU utilization, memory, I/O. Cgroups can have control over the allocation, managing and monitoring of system resources. Hardware resources can also be divided among the tasks and the users which increases efficiency.
Namespaces are the other kernel feature that allows a restricted view of the system. When you login into a Linux machine, you see its file systems, processes, network interfaces. Once you create a container, you enter into a namespace that provides a restricted view and this newly created container has its own file systems, processes and network interface different from the host machine, it’s running on. Each container gets a separate set of kernel resources so that no process can use the resources allocated to other processes.
Is Docker secure?
Most of the newbie developers think that docker is secure and they don’t even do a basic scan of the image which they pulled from the docker hub(docker hub is the world’s largest library and community for container images Browses over 100,000 container images from software vendors, open-source projects, and the community.)Most of the people run randomly Docker images on their system without caring who pushed it or about their authenticity. Docker is not as secure as a virtual machine because a virtual machine does not talk to the host kernel directly. They do not have any access to kernel file systems like /sys and /sys/fs,/proc/*.
It’s simpler to misconfigure docker than it is to misconfigure a virtual machine and the versatility of docker resource sharing opens up further possibilities for both implementation and configuration bugs. However, if you can properly configure the sharing permissions and there are no implementation bugs in Docker or the kernel, Docker offers far more fine-grained sharing than hardware virtualization and can provide you with better overall protection.
In a virtual machine, first the hypervisor would be attacked not the kernel.
Docker security
There are four major areas to consider when reviewing Docker security: the intrinsic security of the kernel and its…
docs.docker.com
Open attack surfaces for Docker containers:
- If the attacker is the one who can start containers (i.e., has access to the Docker API), then he immediately, without further action, has full root access to the host. This is well known for years, has been proven, and is not under debate by anyone (Google or SE immediately give you simple command lines which do not need a particular container to work)
- If the attacker manages to get root inside the container, then you’re in trouble. In effect, he/she can then do pretty arbitrary kernel calls and try to affect the host kernel. Unfortunately, many docker images seem to run their stuff as root and skip the USER in the Dockerfile — this is not a Docker problem but a user problem.
- Vulnerabilities in the kernel Containers on a server share the same kernel as the host, so if the kernel has an exploitable flaw, it could be used to break out of the container and into the host.
- If you have a Bad configuration then you have access to a container that is running with — privileged, you are most likely be able to gain access to the underlying host.
- If you have a container that mounts a host filesystem, you can probably modify things in that filesystem to escalate privileges to the host.
- Mounting the Docker socket inside a container to allow the container to understand the state of the Docker daemon is a relatively common (and dangerous) practice in Docker containers. This gives the host a quick breakout.
Container escape using docker.sock:
- Docker.sock is a unique socket which acts as the backbone for managing your containers, when you type any docker commands using your docker client, your docker client interacts with this docker.sock and manages your container
- By default, when the docker command is executed on a host, an API call to the docker daemon is made via a non-networked UNIX socket located at/var/run/docker.sock.
- This socket file is the main API to control any of the docker containers running on that host.
- However, many containers and guides require you to expose this socket file as a volume within a container or in some cases, expose it on a TCP port.
- Docker containers that expose /var/run/docker.sock, locally or remotely, could lead to a full environment take over.
- Access to /var/run/docker.sock is equivalent to root.
- Now let’s assume that the attacker somehow landed on the container and got a shell on the container where docker.sock is mounted and in this module we’ll see how an attacker can exploit this vulnerability and take over this container.
- First, we need to simulate the environment by creating a container with docker.sock mounted on it, for that I use an alpine image. Let’s start an alpine container and name it as sock using the following command:
$ docker run -itd --name sock -v/var/run/docker.sock:/var/run/docker.sock alpine:latest
- Use the following command to get the shell on the sock container.
$ docker exec -it sock sh
- Now, we have a container with a docker socket mounted on it , lets see how to exploit it. In the first step, install the docker client inside this container and create another container using this container.
$ apk update
$ apk add -u docker
- The following command creates a new container which communicates to the underlying container through Unix socket and -v /:/test:ro mount the root directory of the host to the new container and by using sh it returns the shell of the new container.
$ docker -H unix:///var/run/docker.sock run -it -v /:/test:ro-t alpine sh
- Once we have the shell of the new container, we can go to the /test directory which is the mount point of the root directory of the host.
$ cd /test$ ls
- The output of ls command will list out all the files and directories present in the root directory of host, which implies that we got a foothold of the host machines filesystem.
Exploiting Privileged Container :
- When a container runs with — privileged flag, it gives many capabilities to the container. An attacker who is inside the container can take advantage of these capabilities to escape the container and gain a foothold on the host.
- One of the capabilities provided by — privileged flag is CAP_SYS_MODULE
- Capability, an attacker who has access to the container, with this capability can escape from the container by installing a kernel module directly to the host’s kernel.
- We can use it to get a reverse shell on the host where the privileged container is running.
$ docker run -itd --name priv --privileged alpine
$ docker exec -it priv sh
$ apk add -U libcap
$capsh --print
- Now that we have cap_sys_module capability enabled in our container, we can escape from the container by installing a kernel module directly to the host’s kernel.
Vulnerabilities
- Most of the developers don’t even know about these vulnerabilities.
Now-a-days, docker is widely in use and most of the developers who work with docker think that it is very secure but the reality is different.
- The top 10 official Docker images with more than 10 million downloads each contain at least 30 vulnerabilities.
- Among the top 10 most popular free certified docker images,50 percent have known vulnerabilities.
- 80% of developers don’t test their docker images during development.
- 50% of developers don’t scan their Docker images for vulnerabilities at all.
- 20% of the Docker images with vulnerability could have been solved by a simple image rebuild.
Top 5 Docker Vulnerabilities You Should Know
Offering more than just another buzzword, containers have become one of the biggest trends in software development over…
resources.whitesourcesoftware.com
Some steps for providing better security
- Most base images have vulnerabilities so we should use very less base images because they can have vulnerabilities.
- There are many software to scan images so we should scan our docker images before running it.
- One container for one software/program. If in one container, we install so many software then it will have more vulnerabilities so for fewer vulnerabilities, we should keep one docker container for one software.
- Before pulling an image from the docker hub, we should make sure that the image is pushed by the genuine publisher and check the authenticity of the image and that image can be tempered.
- Don’t use the root user for running docker image. Instead of using the root user, create a dedicated user for running docker image.
- As part of your production pipeline, rebuild your docker files.
- Scan your Docker images as part of your development workflow and in your CI/CD pipelines. Scan your Docker images in your CI/CD pipelines and as part of your development workflow.
- Monitor your Docker containers in production by automatically scanning your base images and packages.
5 tips for securing your Docker containers
Once you start making use of Docker, you'll want to consider the security of your server and containers. Her are five…
Docker Container Security 101: Risks and 33 Best Practices |
Containers, along with orchestrators such as Kubernetes, have ushered in a new era of application development…
Conclusion
In this article we have learned about the vulnerabilities of docker, its type and how to secure our Docker container from it’s vulnerabilities. I have even discussed docker attacks and how do we protect our docker from such attacks and also through this article you all will learn why virtual machines are more secure than docker and in docker, we must do regular scans of images and containers because there are so many vulnerabilities and if not scanned on regular basis then it can cause attacks which will result as a security issue.
References:
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THANKS FOR READING.
ENJOY YOUR CODING! | https://aaryan126.medium.com/how-secure-is-docker-d01154598ed1 | CC-MAIN-2021-49 | refinedweb | 2,242 | 50.87 |
This article describes how integer objects are managed by Python internally.
An integer object in Python is represented internally by the structure PyIntObject. Its value is an attribute of type long.
typedef struct { PyObject_HEAD long ob_ival; } PyIntObject;
To avoid allocating a new integer object each time a new integer object is needed, Python allocates a block of free unused integer objects in advance.
The following structure is used by Python to allocate integer objects, also called PyIntObjects. Once this structure is initialized, the integer objects are ready to be used when new integer values are assigned to objects in a Python script. This structure is called “PyIntBlock” and is defined as:
struct _intblock { struct _intblock *next; PyIntObject objects[N_INTOBJECTS]; }; typedef struct _intblock PyIntBlock;
When a block of integer objects is allocated by Python, the objects have no value assigned to them yet. We call them free integer objects ready to be used. A value will be assigned to the next free object when a new integer value is used in your program. No memory allocation will be required when a free integer object’s value is set so it will be fast.
The integer objects inside the block are linked together back to front using their internal pointer called ob_type. As noted in the source code, this is an abuse of this internal pointer so do not pay too much attention to the name.
Each block of integers contains the number of integer objects which can fit in a block of 1K bytes, about 40 PyIntObject objects on my 64-bit machine. When all the integer objects inside a block are used, a new block is allocated with a new list of integer objects available.
A singly-linked list is used to keep track of the integers blocks allocated. It is called “block_list” internally.
A specific structure is used to refer small integers and share them so access is fast. It is an array of 262 pointers to integer objects. Those integer objects are allocated during initialization in a block of integer objects we saw above. The small integers range is from -5 to 256. Many Python programs spend a lot of time using integers in that range so this is a smart decision.
#define NSMALLPOSINTS 257 #define NSMALLNEGINTS 5 static PyIntObject *small_ints[NSMALLNEGINTS + NSMALLPOSINTS];
The integer object representing the integer -5 is at the offset 0 inside the small integers array. The integers object representing -4 is at offset 1 …
What happens when an integer is defined in a Python script like this one?
>>> a=1 >>> a 1
When you execute the first line, the function PyInt_FromLong is called and its logic is the following:
if integer value in range -5,256: return the integer object pointed by the small integers array at the offset (value + 5). else: if no free integer object available: allocate new block of integer objects set value of the next free integer object in the current block of integers. return integer object
With our example: integer 1 object is pointed by the small integers array at offset: 1+5 = 6. A pointer to this integer object will be returned and the variable “a” will be pointing to that integer object.
Let’s a look at a different example:
>>> a=300 >>> a 300
300 is not in the range of the small integers array so the next free integer object’s value is set to 300.
If you take a look at the file intobject.c in the Python 2.6 source code, you will see a long list of functions taking care of operations like addition, multiplication, conversion… The comparison function looks like this:
static int int_compare(PyIntObject *v, PyIntObject *w) { register long i = v->ob_ival; register long j = w->ob_ival; return (i < j) ? -1 : (i > j) ? 1 : 0; }
The value of an integer object is stored in its ob_ival attribute which is of type long. Each value is placed in a register to optimize access and the comparison is done between those 2 registers. -1 is returned if the integer object pointed by v is less than the one pointed by w. 1 is returned for the opposite and 0 is returned if they are equal.
That’s it for now. I hope you enjoyed the article. Please write a comment if you have any feedback.
interesting post!
i find very fascinating to look at how my favourite language manage data structures.
Link | June 22nd, 2011 at 2:18 pm
Great! Thanks!
Link | June 23rd, 2011 at 8:12 am
Thanks for this and your other posts on Python data structure implementation. I’m teaching an intro CS course this term, using Python, and decided to spend a lecture on “opening up the machine” for my students. These articles are invaluable for my lecture prep.
Link | November 18th, 2012 at 8:43 pm
“The small integers range is from -5 to 257”. Not 257 but 256. Zero is counted as a positive integer.
Link | May 26th, 2013 at 3:04 pm
@Kirill: Thanks! I corrected it.
Link | June 8th, 2013 at 8:02 am
Hi, nice article.
Also, I was linked to this article with a search trying to find out how Python makes up the ability to represent huge integers.
Link | July 24th, 2013 at 9:30 pm
So looking at the PyInt_fromLong logic, I don’t see any way for the virtual machine to reuse integer objects. For example, if I refer to the literal number “300” multiple times in my Python script, will the virtual machine re-create a “300” integer object every time, or will it reuse the first 300 object I created? 300 is not within the span of the “small integer” array.
Link | September 3rd, 2013 at 4:40 pm
@Ricky: In the case of 300, it will use a free integer object from the linked-list each time.
Link | September 5th, 2013 at 7:05 pm
If you do
python -c “a = 1000000; b = 1000000; print a is b; print 1000000 is 1000000”
it prints True twice, which seems to imply that source code literal integers are the same.
Not sure exactly how it works, but I would hope that literals aren’t recreated unnecessarily.
Link | February 27th, 2014 at 7:32 pm
1.
>>> a = 256
>>> b = 256
>>> a is b
True
2.
>>> a = 257
>>> b = 257
>>> a is b
False
3.
>>> a, b = 257, 257
>>> a is b
True
4.
>>> a = 257; b = 257
>>> a is b
False
2. and 4. are same.
but 3. is interesting =)
Link | April 1st, 2015 at 4:36 am
Nicely written article. Gave good understanding of the internal implementation
Link | April 24th, 2015 at 9:32 am
@Dustin That line is interpreted as 1 code object. If you were to do that over multiple lines, then your answer would be different because python would parse multiple lines into multiple code objects. Basically, when python creates a code object from a chunk of syntax, it will scan for all constants and make only 1 object to represent a given constant.
Link | May 23rd, 2015 at 12:36 pm | http://www.laurentluce.com/posts/python-integer-objects-implementation/ | CC-MAIN-2018-17 | refinedweb | 1,188 | 62.48 |
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Detecting CSS Capability
Dr. Evil
02-17-2005, 08:23 AM
Well, this may sound really stupid, but is there a way to detect if a browser has CSS capability? Could you just test in javascript for that_object.style?
ronaldb66
02-17-2005, 10:31 AM
I must admit I really know rather little about JavaScript, but I guess you could detect the availability of certain properties in such a way.
What I'm really curious about is: why would you want to? There's a number of ways to feed adapted style sheets to ancient browsers; if a browser doesn't support any CSS, it'll simply ignore it, and for browsers lacking specific support of certain properties, or misinterpreting them, there are CSS hack and filters.
The whole concept of browser sniffing has proven to be a misconception in numerous cases; I don't think reintroducing it for CSS support sniffing would be such a good plan.
If you really want to pursue the JavaScript approach, though, I think one of the JavaScript forums might yield a better response. Please don't cross-post; I'll request a moderator to move this thread.
Dr. Evil
02-17-2005, 10:57 AM
Ok, thanks. I was just curious because my whole design was practically based on CSS and I didn't know how compatible it would be with older browsers, so I came to the conclusion of detecting CSS capability and displaying the page accordingly. However, this dosn't seem like such a great idea, now that you mention it.
ronaldb66
02-17-2005, 11:57 AM
It's as they always say: "code for standards, not for browsers".
This doesn't mean that you don't have to make some effort to acommodate browsers with partial or no CSS support, but if the page's structure is solid and the semantics are firmly in place, the page should still be usable even without CSS, albeit not that pretty.
Also, don't forget that depending on JavaScript makes this detection fail for a certain percentage of browsers that don't support JavaScript or have it disabled or blocked in some way. Again, not the end of the world, but it shouldn't hinder the site's usability.
Okay, I'll get off my soap box now... :o
brothercake
02-17-2005, 11:58 AM
I can see where you're coming from, but consider - what would you do different if CSS were not supported?
ronaldb66
02-17-2005, 12:12 PM
Who- me? Or Dr. Evil?
brothercake
02-17-2005, 12:52 PM
Dr. Evil - it was just a general throwback to consider - "what would you actually do if you could do this"
Should have used a [quote] ...
ronaldb66
02-17-2005, 01:50 PM
Ah.
Good question, though.
Dr. Evil
02-17-2005, 06:50 PM
I can see where you're coming from, but consider - what would you do different if CSS were not supported?.
Puffin the Erb
02-17-2005, 07:16 PM
I agree with what Ronald said.
I would add.....
Make sure your pages 'degrade gracefully', they may look different in different browsers but the important thing is that they are readable, searchable and navigatable - don't rely on JavaScript or CSS for these aspects of the site.
Keep in mind the formatting effects associated with HTML tags - don't override them with CSS, for example making a <div> an inline tag with display:inline.
Make sure your pages are tested in a browser in which script and css are not supported ( Lynx, Links, Elinks, Dillo etc) .
Text browsers do not support css or client-side scripting but they render tags. A text browser user does not expect, or need, their browser to render fancy effects they want to be able to read, search and navigate content.
Avoid creating separate versions of your pages - this will make the site very difficult to maintain - just mark your content up logically.
Ronald can have the soap box back now. :)
ReadMe.txt
02-17-2005, 08:28 PM.
If you make the JS set them as hidden in the first place the script will degrade gracefully.
brothercake
02-17-2005, 09:22 PM
Still, there could be times where it's desirable - a scripted, list-based menu for example - in the unlikely but possible circumstance of a client who has CSS turned off but scripting turned on, what they'd get is an unstyled list with dynamic open/close behaviors - and that's not good at all; it would much better in that situation if one could, after all, use scripting to detect whether CSS is available.
So ... with this discussion as proviso - of when it's not appropriate, and how to degrade naturally through good separation of content, logic and presentation ... here's how to detect whether CSS is enabled:
Some elements have default properties which are predictable in all browsers. For example, links always have inline display. So if you make a specific link with block-level display, then read its display property out of currentStyle or getComputedStyle - if the value is not "inline" you can deem CSS to be enabled; otherwise you can deem that it isn't.
It won't work in Safari or early Opera 7 builds, but it works for later opera, moz and ie:
//check that nav styles are enabled
function cssIsEnabled()
{
//reference to test link
var linkObj = document.getElementById("testLink");
//link display
var linkDisplay = "inline";
//retrieve the computedStyle display of the link
//this is mozilla and opera 7.2+
if(typeof document.defaultView != "undefined" && typeof document.defaultView.getComputedStyle != "undefined")
{
linkDisplay = document.defaultView.getComputedStyle(linkObj,"").getPropertyValue("display");
}
//retrieve the currentStyle display of the link
//this is internet explorer
else if(typeof linkObj.currentStyle != "undefined")
{
linkDisplay = linkObj.currentStyle.display;
}
//if display is not 'inline' then CSS is deemed to be enabled
//this is also the fallback position if no property is detected
//which is Safari and Opera <7.2
return (linkDisplay != "inline");
};
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Improving the performance of Visual Studio
I first heard about the idea a couple of weeks ago, when I was looped into a discussion between Craig Neable and the Channel 9 team. Craig and his boss Carter Maslan work every day with developers who are trying to sell their bosses on the benefits of .NET. I figure this normally means a lot of “push” marketing – preparing datasheets and case studies and competitive analyses. Then Carter had the slightly-crazy idea that the developers themselves know why they want Visual Studio 2005, so why not give them the floor? Or, in his own words:?
Craig bounced a few namespace ideas off me, some of which I teased him for unmercifully (sorry Craig!), and then we got down to playing with prototypes. To kick things off, Craig and Carter have added the reasons they think companies will love Visual Studio 2005, but now the High Five Wikisheets are open to developers everywhere. Got a great reason that convinced your boss to switch? They want to hear it. Convinced that Visual Studio 2005 adds no value? They want to hear that too. Spotted a connection between two different scenarios? Refactor the wiki — or just join in the conversation on Channel 9!
Now, I wonder what Rick Segal and Seth Godin think?
Technorati tags: Cluetrain, Marketing, Microsoft, Wikis | http://blogs.msdn.com/jonathanh/archive/2005/08/08/448999.aspx | crawl-002 | refinedweb | 224 | 74.29 |
dinsdag 2 september 2008
Google Chrome
I love their new 'start-page' concept, for instance.
They've also created a comic where they explain the concepts and techniques of Chrome. Very interesting as well. :)
maandag 25 augustus 2008
On reading books ....
D...
dinsdag 19 augustus 2008
Locking system with aspect oriented programming
Intro
A few months ago, I had to implement a 'locking system' at work.
I will not elaborate to much on this system, but it's intention is that users can prevent that certain properties of certain entities are updated automatically;
The software-system in where I had to implement this functionality, keeps a large database up-to-date by processing and importing lots of data-files that we receive from external sources.
Because of that, in certain circumstances, users want to avoid that data that they've manually changed or corrected, gets overwritten with wrong information next time a file is processed.
The application where I'm talking about, makes heavy use of DataSets and I've been able to create a rather elegant solution for it.
At the same time, I've also been thinking on how I could solve this same problem in a system that is built around POCO's instead of Datasets, and that's what this post will be all about. :)
Enter Aspects
When the idea of implementing such a system first crossed my mind, I already realized that Aspects Oriented Programming could be very helpfull to solve this problem.
A while ago, I already played with Aspect Oriented Programming using Spring.NET.
AOP was very nice and interesing, but I found the runtime-weaving a big drawback. Making use of runtime weaving meant that you could not directly create an instance using it's constructor.
So, instead of:
MyClass c = new MyClass();you had to instantiate instances via a proxyfactory:
ProxyFactory f = new ProxyFactory (new TestClass());
f.AddAdvice (new MethodInvocationLoggingAdvice());
ITest t = (ITest)f.GetProxy();
I am sure that you agree that this is quite a hassle, just to create a simple instance. (Yes, I know, offcourse you can make abstraction of this by making use of a Factory...).
Recently however, I bumped at an article on Patrick De Boeck's weblog, where he was talking about PostSharp.
PostSharp is an aspect weaver for .NET which weaves at compile-time!
This means that the drawback that I just described when you make use of runtime-weaving has disappeared.
So, I no longer had excuses to start implementing a similar locking system for POCO's.
Bring it on
I like the idea of Test-Driven-Development, so I started out with writing a first simple test:
The advantage of writing your test first, is that you start thinking on how the interface of our class should look like.
This first test tells us that our class should have a
Lock and an
IsLocked method.
The purpose of the
Lock method is to put a 'lock' on a certain property, so that we can avoid that this property is modified at run-time.
The
IsLocked method is there to inform us whether a property is locked or not.
To define this contract, I've created an interface
ILockable which contains these 2 methods.
In order to get this first test working, I've created an abstract class
LockableEntity which inherits from one of my base entity-classes implements this interface.
This
LockableEntity class looks like this:
This is not sufficient to get a green bar on my first test, since I still need an
AuditablePerson class:
These pieces of code are sufficient to make my first test pass, so I continued with writing a second test:
As you can see, in this test-case I define that it should be possible to unlock a property. Unlocking a property means that the value of that property can be modified by the user at runtime.
To implement this simple functionality, it was sufficient to just add an
UnLock method to the
LockableEntity class:
Simple, but now, a more challenging feature is coming up.
Now, we can already 'lock' and 'unlock' properties, but there is nothing that really prevents us from changing a locked property.
It's about time to tackle this problem and therefore, I've written the following test:
Running this test obviously gives a red bar, since we haven't implemented any logic yet.
The most simple way to implement this functionality, would be to check in the setter of the
Name property whether there exists a lock on this property or not.
If a lock exists, we should not change the value of the property, otherwise we allow the change.
I think that this is a fine opportunity to use aspects.
Creating the Lockable Aspect
As I've mentionned earlier, I have used PostSharp to create the aspects. Once you've downloaded and installed PostSharp, you can create an aspect rather easy.
There is plenty of documentation to be found on the PostSharp site, so I'm not going to elaborate here on the 'getting started' aspect (no pun intended).
Instead, I'll directly dive into the
Lockable aspect that I've created.
This is how the definition of the class that defines the aspect looks like:
Perhaps I should first elaborate a bit on how I would like to use this
Lockable aspect.
I'd like to be able to decorate the properties of a class that should be 'lockable' with an attribute. Like this:
Decorating a property with the
Lockable attribute, means that the user should be able to 'lock' this property. That is, prevent that it gets changed after it has been locked.
To be able to implement this, I've created a class which inherits from the OnMethodInvocationAspect class (which eventually inherits from
Attribute).
Why did I choose this class to inherit from?
Well, because there exists no
OnPropertyInvocation class or whatsoever.
As you probably know, the getters and setters of a property are actually implemented as
get_ and
set_ methods, so it is perfectly possible to use the
OnMethodInvocationAspect class to add extra 'concerns' to the property.
This extra functionality is written in the
OnInvocation method that I've overriden in the
LockableAttribute class.
In fact, it does nothing more then checking whether we're in the setter method of the property, and if we are, check whether there exists a lock on the property.
If there exists a lock, we won't allow the property-value to be changed. Otherwise, we just make sure that the implementation of the property itself is called.
The implementation looks like this:
Here, you can see that we use reflection to determine whether we're in the setter-method or in the getter-method of the property; we're only interested if this property is locked if we're about to change the value of the property.
Next, we need to get the name of the property for which we're entering the setter method. This is done via the
GetPropertyForSetterMethod method which uses reflection as well to get the
PropertyInfo object for the given setter-method.
Once this has been done, I can use the
IsLocked method to check whether this property is locked or not.
Note that I haven't checked whether the conversion from
eventArgs.Delegate.Target to
ILockable has succeeded or not. More on that later ...
When the property is locked, I call the
OnAttemptToModifyLockedProperty method (which is declared in
ILockable), and which just raises the
LockedPropertyChangeAttempt event (also declared in the
ILockable interface). By doing so, the programmer can decide what should happen when someone / something attempts to change a locked property. This gives a bit more control to the programmer and is much more flexible then throwing an exception.
When the property is not locked, we let the setter-method execute.
With the creation of this aspect, our third test finally gives a green bar.
Compile time Validation
As I've said a bit earlier, I haven't checked in the
OnInvocation method whether the Target really implemented the
ILockable interface before I called methods of the
ILockable type.
The reason for this , is quite simple: the
OnMethodInvocationAspect class has a method
CompileTimeValidate which you can override to add compile-time validation logic (hm, obvious).
I made use of this to check whether the types where I've applied the
Lockable attribute really are
ILockable types:
Note that it should be possible to make this code more concise, but I could not just call
method.DeclaringType.GetInterface("ILockable")since that gave a
NotImplementedExceptionwhile compiling. Strange, but true
Now, when I use the
Lockable attribute on a type which is not
ILockable, I'll get the following compiler errors:
Pretty neat, huh ?
Now, what's left is a way to persist the locks in a datastore, but that will be a story for some other time ...
maandag 28 juli 2008
NHibernate in a remoting / WCF scenario
This ?
zaterdag 5 juli 2008
NHibernate IInterceptor: an AuditInterceptor ...
dinsdag 1 juli 2008
NHibernate Session Management
I_10<<.
maandag 30 juni 2008
New Layout
I've changed the layout of my weblog, I hope you like it.
If you have any remarks regarding the layout, if you don't find it readable, if you miss something, please let me know.
vrijdag 13 juni 2008
Setting Up Continuous IntegrationPart II: configuring CruiseControl.NET
Now that we've created our buildscript in part I, it's time to set up the configuration file for CruiseControl.NET
The ccnet.config file and multiple project-configurations :)
MSBuild doesn't support my sln file format.
The MSBuild XmlLogger Issue.
zaterdag 7 juni 2008
Setting Up a Continuous Integration process using CruiseControl.NET and MSBuild.
Part I: creating the MSBuild build script
Intro. :)
Requirements:
- Make sure that the latest buildscript will be used
- Clean the source directory
- Get the latest version of the codebase out of Visual SourceSafe
- Build the entire codebase
- Execute the unit tests that I have using NUnit
- Perform a statical code analysis using FxCop
The MSBuild build-script.
Skeleton of the buildscript'.
Clean Target:
Getlatest TargetIn order to get the latest version of the source out of SourceSafe, I’ve created the following step:
Here, I just make use of the VssGet Task that is part of the MSBuild Community Tasks project.
Also, notice that this Target depends on the createdirs Target; this means that, when you execute the getlatest Target, the createdirs Target will be executed before the getlatest Target is executed.
The createdirs tasks looks like this:
BuildAll Target.
NUnit Target).
FxCop Target ...
Executing Targets via MSBuild.
zondag 20 april 2008
using directives within namespaces
Sometimes, I come across code examples where the programmer puts his using directives within the namespace declaration, like this:
namespace MyNamespace
{
using System;
using System.Data;
using SomeOtherNamespace;
public class MyClass
{
}
}
I am used to put my using directives outside the namespace block (which is no surprise, since VS.NET places them by default outside the namespace declaration when you create a new class):
using System;
using System.Data;
namespace MyNamespace
{
public class MyClass
{
}
}
So, I'm wondering: what are the advantages of placing the using directives within the namespace declaration ?
I've googled a little bit, but I haven't found any clue why I should do it as well. Maybe you'll know a good reason, and can convice me to adapt my VS.NET templates ?
woensdag 12 maart 2008
VS.NET 2008: Form designer not working on Windows Vista
I_21<<
maandag 28 januari 2008
Debugging the .NET framework
dinsdag 15 januari 2008
Cannot open log for source {0} on Windows 2003 Server
I
| http://fgheysels.blogspot.com/2008/ | CC-MAIN-2017-30 | refinedweb | 1,941 | 60.75 |
Summary
Resizes a raster by the specified x and y scale factors.
Usage
The output size is multiplied by the scale factor for both the x and y directions. The number of columns and rows stays the same in this process, but the cell size is multiplied by the scale factor.
The scale factor must be positive.
A scale factor greater than one means the image will be rescaled to a larger dimension, resulting in a larger extent because of a larger cell size.
A scale factor less than one means the image will be rescaled to a smaller dimension, resulting in a smaller extent because of a smaller cell size.
You can save the output to BIL, BIP, BMP, BSQ, DAT, Esri Grid, GIF, IMG, JPEG, JPEG 2000, PNG, TIFF, MRF, or CRF format, or any geodatabase raster dataset.
When storing a raster dataset to a JPEG format file, a JPEG 2000 format file, or a geodatabase, you can specify a Compression Type value and a Compression Quality value in the geoprocessing environments.
This tool supports multidimensional raster data. To run the tool on each slice in the multidimensional raster and generate a multidimensional raster output, be sure to save the output to CRF.
Supported input multidimensional dataset types include multidimensional raster layer, mosaic dataset, image service, and CRF.
Parameters
arcpy.management.Rescale(in_raster, out_raster, x_scale, y_scale)
Code sample
This is a Python sample for the Rescale tool.
import arcpy arcpy.Rescale_management("c:/data/image.tif", "c:/output/rescale.tif", "4", "4")
This is a Python script sample for the Rescale tool.
##==================================== ##Rescale ##Usage: Usage: Rescale_management in_raster out_raster x_scale y_scale import arcpy arcpy.env.workspace = r"C:/Workspace" ##Rescale a TIFF image by a factor of 4 in both directions arcpy.Rescale_management("image.tif", "rescale.tif", "4", "4")
Environments
Licensing information
- Basic: Yes
- Standard: Yes
- Advanced: Yes | https://pro.arcgis.com/en/pro-app/latest/tool-reference/data-management/rescale.htm | CC-MAIN-2022-27 | refinedweb | 308 | 54.22 |
RubyMine 2020.2 EAP2: Improvements for Liquid Template Language, New Intention Actions, and More
RubyMine 2020.2 EAP2 is now available! In this version, we’re continuing our work on Liquid template language support. We’ve also added new intention actions, and improved the folding of
if/while/for statements:
- Liquid template language support improvements
- New intention actions
- Improved readability of folded if/while/for statements
Liquid template language support improvements
Live templates
This EAP version comes with live templates for Liquid. Live templates (or code snippets) allow you to insert frequently-used constructions into your code. These can be conditions, blocks, loops, and so on.
To invoke a live template, start typing and press Tab:
To see and configure the available live templates, go to the Settings/Preferences | Editor | Live Templates page:
Folding
RubyMine now recognizes code blocks and allows you to fold tags with code blocks, for example the
{% liquid %} tag, if-else statements, and so on:
Reformatting the code
RubyMine lets you reformat your Liquid code according to the requirements you’ve specified in the Code Style settings. Access the settings in Settings/Preferences | Editor | Code Style.
To run reformatting, either go to Code | Reformat or select the code fragment you want to reformat and press ⌥⌘L (Ctrl+Shift+L) in the editor. You can also reformat a file or a group of files. Learn more about it in the documentation.
We’ve already implemented support for the
{% comment %}...{% endcomment %} tags. Starting with this build, comments inside Liquid tags are also supported:
New intention actions
Merge/split sequential ‘if’s
This intention action is available on the
elsif or
if keywords. It suggests merging two branches if the code inside these branches is exactly the same. RubyMine will then combine the two conditions using an || operator and add parenthesis if necessary.
The Split into multiple ‘if’s action will do the opposite.
Expand or flatten namespace
These intention actions will expand lines with the
:: scope resolution operators into nested modules and vise versa. Note that the Flatten namespace operator will flatten the modules above it:
Improved readability of folded if/while/for statements
For folded constructs, like
if,
while, and
for, RubyMine will now show the condition or variables used in the first line. This approach makes it easier to see whether you need to unfold the construct and look into it.
2
Happy Developing!
The RubyMine team | https://blog.jetbrains.com/ruby/2020/06/rubymine-2020-2-eap2/ | CC-MAIN-2020-29 | refinedweb | 401 | 53.1 |
Opened 9 years ago
Closed 8 years ago
Last modified 8 years ago
#5858 closed defect (fixed)
Code inserted is not XHTML valid
Description
Code inserted is not XHTML valid. Around
<label> should be a
<div>. Fix is simple, just add
tag.div(...) around result in
_user_field_input:
def _user_field_input(self, req): return tag.div(tag.label("Filter users: ", tag.input(type="text", name="users", value=req.args.get('users', ''), style_="width:60%")))
Attachments (0)
Change History (2)
comment:1 Changed 8 years ago by
comment:2 Changed 8 years ago by
(In [7595]) Patch from asic_druide to fix #5400, #5421 and #5858. Thanks!
Note: See TracTickets for help on using tickets.
Attached version fixes #5400, #5421 and #5858. | https://trac-hacks.org/ticket/5858 | CC-MAIN-2018-30 | refinedweb | 118 | 68.67 |
I stumbled across a great resource awhile back at. You can find a large assortment of assets for game development including art and even music, and as the name implies, you are free to use most of it in your projects. This coupled with the enjoyment I had with the random world creator I made in the last post and I have decided to move away from the purely text based RPG. It will be easier than I thought to add some graphics, so in this post, I will show how you could extend the Procedural World Visualizer into a sprite based equivalent.
Prerequisites
If you haven’t already read my previous post, you should do so now. It includes and explains some source code you will need to complete this demo. I won’t be re-explaining any of that portion of the code, and will merely comment on how to use it to drive the sprites that get picked to layout a level.
You will also need art resources. The sprite pack I ended up choosing is found here Crawl Tiles – a 2.7 Mb zip file full of TONS of sprites for terrain, dungeons, decoration, monsters, and probably everything else you could hope to implement in our simple RPG project. Of course you could use any sprites you like or create your own.
Create the Scene
The first thing we need is a way to show our world on screen. We will do this by taking advantage of Unity’s Grid Layout Group component to show our sprite images.
- Create a new scene.
- Add a Panel. From the menubar, choose “GameObject->UI->Panel”. I named my panel “Scenery” and removed the default Image and Canvas Renderer components.
- Modify the Scenery Panel’s Rect Transform so that all anchor and pivot fields use a value of 0.5. Then zero out the position, and set the size to width:640 and height:480.
- Add the component, “Grid Layout Group” to the Panel. Set the “Cell Size” to 32 on both X and Y. Also set the “Start Corner” and “Child Alignment” properties to “Lower Left”.
- Create a new script called “PerlinSpriteVisualizer.cs” and add it to the panel.
- Create a new Image object. From the menubar, choose “GameObject->UI->Image”.
- Create a prefab from the new Image object by dragging it into the Project pane. Then delete the object from the scene. The prefab we just created will be assigned to a property reference on the PerlinSpriteVisualizer class in a moment.
- Select the parent canvas object. In the inspector, modify the Canvas Scaler’s “Ui Scale Mode” to “Scale With Screen Size”. Put in a “Reference Resolution” of “x:640 y:480”. Set the “Screen Match Mode” to “Shrink”. All of these settings will help to make sure that our interface stuff will scale to fill the camera screen regardless of the aspect ratio and resolution we actually end up using.
Get Some Sprites On Screen
We are going to need a class which can take an x and y position (the tile’s coordinates on screen) and determine what sprite should appear. This functionality is defined in the abstract base class:
public abstract class SpritePicker : MonoBehaviour { public abstract Sprite Pick (int x, int y); }
Our first few concrete subclasses are almost identical to the idea I used for the EnemySpawner post – one Fixed that always returns the same sprite, one Random that can return any sprite from an array at random, and one weighted, which is more likely to return certain sprites based on higher weights. If you find this or the topic of polymorphism confusing, you may want to go back and read that post.
public class FixedSpritePicker : SpritePicker { public Sprite sprite; public override Sprite Pick (int x, int y) { return sprite; } }
public class RandomSpritePicker : SpritePicker { public Sprite[] sprites; public bool persistant = true; public override Sprite Pick (int x, int y) { if (persistant) return PickPersistantRandom(x, y); else return PickRandom(x, y); } Sprite PickPersistantRandom (int x, int y) { // Store the seed that had been active int oldSeed = UnityEngine.Random.seed; // Seed the random generator based on the coordinates UnityEngine.Random.seed = (x + 3) ^ (y + 7); // Pick a sprite to return Sprite retValue = sprites[ UnityEngine.Random.Range(0, sprites.Length) ]; // Restore the original seed UnityEngine.Random.seed = oldSeed; return retValue; } Sprite PickRandom (int x, int y) { return sprites[ UnityEngine.Random.Range(0, sprites.Length) ]; } }
[System.Serializable] public class WeightedSprite { public Sprite sprite; public int weight; } public class WeightedRandomSpritePicker : SpritePicker { public WeightedSprite[] options; public bool persistant = true; int optionCount; int wheelSize; void Start () { BuildWheel(); } public void BuildWheel () { wheelSize = 0; optionCount = options.Length; for (int i = 0; i < optionCount; ++i) { wheelSize += options[i].weight; } } public override Sprite Pick (int x, int y) { if (persistant) return PickPersistantRandom(x, y); else return SpinWheel(); } Sprite PickPersistantRandom (int x, int y) { // Store the seed that had been active int oldSeed = UnityEngine.Random.seed; // Seed the random generator based on the coordinates UnityEngine.Random.seed = (x + 3) ^ (y + 7); // Pick a sprite to return Sprite retValue = SpinWheel(); // Restore the original seed UnityEngine.Random.seed = oldSeed; return retValue; } Sprite SpinWheel () { int roll = UnityEngine.Random.Range(0, wheelSize); int sum = 0; for (int i = 0; i < optionCount; ++i) { WeightedSprite option = options[i]; sum += option.weight; if (sum >= roll) return option.sprite; } return null; } }
Note that the Random and Weighted versions of the Sprite Picker both include a boolean indicating whether or not the selection is “persistant” which I default to true. Normally, every time you choose a random value, Unity returns a different value. However, because we are using the random value to pick a sprite to appear in our wold, we want to make sure that the sprite that was picked for any particular location is always the same. For example, imagine you have your character move right, scrolling the background by one tile. Then you move back to where you were. If the tiles which loaded randomly were not persistent, you might notice flowers appearing or disappearing in the grass.
The way that this randomness is persisted is by assignment of a “seed” value to the Random class. The value that Random actually returns is not actually random, but because the seed changes, the result we get back does. By specifying a seed, the value we generate will always be consistent.
Next implement the PerlinSpriteVisualizer script which we created earlier. Assign the Image prefab I had you create to the “tilePrefab” property.
public class PerlinSpriteVisualizer : MonoBehaviour { #region Properties public GameObject tilePrefab; public SpritePicker spritePicker; public Image[] tiles; public int xPos, yPos; int _w, _h; #endregion #region MonoBehaviour void Start () { Init (); Refresh(); } void Update () { bool dirty = false; if (Input.GetKeyUp(KeyCode.UpArrow)) { yPos++; dirty = true; } if (Input.GetKeyUp(KeyCode.DownArrow)) { yPos--; dirty = true; } if (Input.GetKeyUp(KeyCode.LeftArrow)) { xPos--; dirty = true; } if (Input.GetKeyUp(KeyCode.RightArrow)) { xPos++; dirty = true; } if (dirty) Refresh(); } #endregion #region Public public void Refresh () { for (int y = 0; y < _h; ++y) { for (int x = 0; x < _w; ++x) { int index = y * _w + x; tiles[index].sprite = spritePicker.Pick(x + xPos, y + yPos); } } } #endregion #region Private void Init () { RectTransform rt = GetComponent<RectTransform>(); GridLayoutGroup glg = GetComponent<GridLayoutGroup>(); _w = Mathf.FloorToInt(rt.rect.width / glg.cellSize.x); _h = Mathf.FloorToInt(rt.rect.height / glg.cellSize.y); int count = _w * _h; tiles = new Image[ count ]; for (int i = 0; i < count; ++i) tiles[i] = CreateTile(); } Image CreateTile () { GameObject instance = Instantiate(tilePrefab) as GameObject; instance.transform.SetParent(transform, false); return instance.GetComponent<Image>(); } #endregion }
Create an Empty GameObject in the scene called “Sprite Pickers”. This will be an object I use for organization in the scene hierarchy. Create another Empty GameObject in the scene called “Fixed Sprite Picker” and parent it to the “Sprite Pickers” object. Attach the “FixedSpritePicker” script as a component and assign any of your project’s sprites to its “Sprite” property – I used a tileable dirt sprite – “grass_full” (poorly named considering it is a picture of dirt) if you are using the same texture pack.
Assign the “Perlin Sprite Visualizer” script’s “Sprite Picker” property to the Fixed sprite picker you just created. If you run the scene now, you should see an array of those images fill the panel without any gaps.
Create two more objects in your scene for the “RandomSpritePicker” and “WeightedSpritePicker” classes just like we did for the “FixedSpritePicker”. Make sure to parent them to the “Sprite Pickers” object to help keep the scene organized. Find a group of sprites that are intended to be swappable, such as versions of grass with and without flowers, weeds, etc. and assign them to the other pickers. Then try running the scene with one of them set as the PerlinSpriteVisualizer’s SpritePicker. The ground should look a lot better when it picks from a variety of sprites over using the exact same tiled sprite. It will help to break up the obvious tiling of the images you are using.
With the scene in play mode, note that you can use the arrow keys to scroll the background. It almost feels like you are exploring a game world already!
Use Some Perlin Noise
Getting sprites on screen was sort of exciting, but I really want to see the Perlin noise examples from the previous post driving the layout of our sprites. Let’s do that now.
Create a new class called “Painter” that will function kind of like the PerlinNoiseVisualizer from the previous post. It takes an x and y position and returns a Color.
public abstract class Painter : MonoBehaviour { public abstract Color Render (int x, int y); }
The concrete implementation presented below works off of the RenderLayer implementation we used in the previous post.
public class RenderLayerPainter : Painter { public RenderLayer[] layers; public Color Render (int x, int y) { Color c = Color.black; for (int i = 0; i < layers.Length; ++i) { c += layers[i].Render(x, y); } c.r = Mathf.Clamp01(c.r); c.g = Mathf.Clamp01(c.g); c.b = Mathf.Clamp01(c.b); c.a = Mathf.Clamp01(c.a); return c; } }
You could create other concrete implementations of Painter, such as one that simply reads an image and passes along its colors. This way you could hand paint levels in photoshop, etc. At the moment that will be left as an exercise for the reader.
Now I will create a SpritePicker subclass which uses a painter to drive the sprite selection. Note that it maps pixel colors to other sprite pickers instead of directly to sprites. This allows a nice hierarchy of complexity, such as saying any white pixel will be dirt, and black will be grass, but if the pickers are implemented as RandomSpritePickers, then the white pixels will be random dirt sprites instead of a constantly repeated dirt sprite.
[Serializable] public class PaletteTheme { public Color color; public SpritePicker picker; } public class PaletteSpritePicker : SpritePicker { public PaletteTheme[] theme; public Painter painter; Dictionary<Color, SpritePicker> map = new Dictionary<Color, SpritePicker>(); void Start () { for (int i = theme.Length - 1; i >= 0; --i) map.Add(theme[i].color, theme[i].picker); } public override Sprite Pick (int x, int y) { Color c = painter.Render (x, y); SpritePicker sp = map.ContainsKey(c) ? map[c] : theme[0].picker; return sp.Pick(x, y); } }
Create another empty GameObject for our Palette Sprite Picker class. Name it and parent it as we have done before and assign it as the Sprite Picker for the Perlin Sprite Visualizer script. I set the Theme to have two entries, the first with a color of black (make sure the alpha is full white) and a picker set to the Random Sprite Picker we created earlier (to assign a random grass tile). The second theme element has a color of white and maps to the Fixed sprite picker (dirt).
Make another empty GameObject called “Painter” and add the “RenderLayerPainter” script. I set it up like the last example of the previous post where there are two Render Layers, the first driven by a “PerlinMapper” at a scale of 0.1 and using a “BarNode” with a value of 0.6. The second layer uses “PerlinMapper” at a scale of 0.15 and a “BandNode” with values of 0.4-0.6 set as the range. These scripts were all defined in the previous post in case you missed them. Assign this painter object to the painter property of the “PaletteSpritePicker” and then run the scene.
You should see a very nice perlin noise driven example. Feel free to explore the world using the arrow keys and notice how the pattern is unique for as long as you explore, but the same coordinates always show the same thing. You can use this fact to hand place certain elements like cities or dungeons if you were so inclined.
Add Transitions
Many tile engines use transition tiles between to ease the visual contrast between two types of environment tiles. The sprite bundle I downloaded has sprites for this purpose: grass_nw, grass_n, grass_ne, etc. It is not hard to create a sprite picker which can evaluate the position of a pixel and its surrounding pixels to determine which transition tile should be added.
We will create a new script called “TransitionSpritePicker” for this purpose.
[Flags] public enum TransitionEdge { None = 0, Top = 1 << 0, Bottom = 1 << 1, Right = 1 << 2, Left = 1 << 3 } public class TransitionSpritePicker : SpritePicker { public Painter painter; public SpritePicker none; public SpritePicker topLeft; public SpritePicker topMiddle; public SpritePicker topRight; public SpritePicker left; public SpritePicker full; public SpritePicker right; public SpritePicker bottomLeft; public SpritePicker bottomCenter; public SpritePicker bottomRight; Dictionary<TransitionEdge, SpritePicker> map = new Dictionary<TransitionEdge, SpritePicker>(); void Start () { map.Add(TransitionEdge.None, none); map.Add(TransitionEdge.Bottom | TransitionEdge.Right, topLeft); map.Add(TransitionEdge.Bottom | TransitionEdge.Left | TransitionEdge.Right, topMiddle); map.Add(TransitionEdge.Bottom | TransitionEdge.Left, topRight); map.Add(TransitionEdge.Top | TransitionEdge.Bottom | TransitionEdge.Right, left); map.Add(TransitionEdge.Top | TransitionEdge.Bottom | TransitionEdge.Left | TransitionEdge.Right, full); map.Add(TransitionEdge.Top | TransitionEdge.Bottom | TransitionEdge.Left, right); map.Add(TransitionEdge.Top | TransitionEdge.Right, bottomLeft); map.Add(TransitionEdge.Top | TransitionEdge.Left | TransitionEdge.Right, bottomCenter); map.Add(TransitionEdge.Top | TransitionEdge.Left, bottomRight); } public override Sprite Pick (int x, int y) { TransitionEdge edge = GetEdges(x, y); SpritePicker sp = map.ContainsKey(edge) ? map[edge] : none; return sp.Pick(x, y); } TransitionEdge GetEdges (int x, int y) { Color myColor = painter.Render (x, y); TransitionEdge edge = TransitionEdge.None; if (myColor == Color.black) return edge; if (painter.Render(x, y+1) == myColor) edge |= TransitionEdge.Top; if (painter.Render(x, y-1) == myColor) edge |= TransitionEdge.Bottom; if (painter.Render(x+1, y) == myColor) edge |= TransitionEdge.Right; if (painter.Render(x-1, y) == myColor) edge |= TransitionEdge.Left; return edge; } }
This example is not as efficient as it could be, because the painter it uses recalculates the color of a location every time it is queried instead of caching values. And in this case, every tile queries itself and the four tiles around it, so the total number of queries grows very quickly. When the math is quick enough, this is not a serious issue, but depending on the amount of tiles and the complexity of the noise, it could become one.
Create another SpritePicker container in your scene for the TransitionSpritePicker script. Use the same Painter we used for the PaletteSpritePicker. I created FixedSpritePickers for all of the transition tiles, and assigned them to the appropriate properties, but the “none” sprite picker was assigned to the random grass picker we already had created. Run the scene and you will see a version similar to the palette version, but with nice transitions – in most of the locations. The problem is that our painter is creating patterns where a transition is required that we didn’t define. For example, a tile that has matches at its top and bottom but neither side. I defaulted to the “none” picker so it shows grass in those locations, but you will probably notice them because the transition edge is completely sharp wherever they appear. Remedies for this issue could include creating the additional transition tile possibilities and adding them to the script, only using this picker on hand painted maps, or caching and evaluating the pattern in the painter to remove / repaint tiles that are not supported. | http://theliquidfire.com/2014/12/18/world-implementation/ | CC-MAIN-2017-51 | refinedweb | 2,696 | 55.54 |
Mike, Matt and I had a blast out at the Bay Area Maker Faire in San Mateo last week. It was great to put some faces to names, and talk to fellow hackers in person.
I had a couple folks asking about their Arduino greenhouse projects, and they mentioned it might be useful to have a humidity sensor to make sure the room didn’t get too dry.
So Paul and Chris built the Humidity Sensor, which is really a Humidity and Temperature Sensor, since it provides an integrated temperature reading as well. I’ve uploaded it over at the Liquidware and Modern Device shop pages.
The Humidity (and Temperature) Sensor is a board that carries the SHT21 digital humidity and temperature sensor from Sensirion. It has 4 pins that can mount directly on the Arduino’s analog pins. Two of the four pins are Ground and +5V; the other two are clock and data pins making up the 2-wire serial interface.
It transmits data to the Arduino over an I2C protocol, and comes with built-in 5V tolerance to be powered directly from the analog pins. The Antipasto Arduino IDE has an integrated library that makes the humidity sensor plug-and-play.
Here’s a quick demo that Will and I put together, which maps out the relative humidity and temperature as a bargraph on the TouchShield Slide. Because the Slide doesn’t occupy any of the Arduino Duemilanove’s analog pins, I was able to mount both the Humidity Sensor and the Slide directly onto the Arduino.
I took a video of the Humidity Sensor bargraphs in action:
And the code is below:
//Humidity sensor bar graphs - Arduino Code
#include <Wire.h>
#include <LibHumidity.h>
#include <HardwareSensor.h>
int h;
int t;
LibHumidity humidity = LibHumidity(0);
void setup() {
Sensor.begin(19200);
}
void loop() {
h = (int)humidity.GetHumidity();
t = (int)humidity.GetTemperature();
Sensor.print("humidity", h);
Sensor.print("temp", t);
delay(5);
}
________________________________________________
//Humidity sensor bar graphs - TouchShield Code
#include <HardwareSensor.h>
int newT;
void setup() {
Sensor.begin(19200);
background(0);
line(0, 120, 320, 120);
text("RHumidity:", 0, 50);
text("Fahrenheit:", 0, 140);
}
void loop() {
if (Sensor.available()) {
int h;
int t;
if (!strcmp(Sensor.getName(), "humidity")) {
h = Sensor.read();
fill(0);
text(h, 230, 80);
delay(10);
fill(0);
rect(0, 70, 200, 30);
{
fill(0, 0, 255);
rect(0, 70, h*2, 30);
}
}
if (!strcmp(Sensor.getName(), "temp")) {
t = Sensor.read();
//convert celcius to fahrenheit
newT = (t*1.8 + 32);
fill(0);
text(newT, 230, 170);
delay(10);
fill(0);
rect(0, 160, 200, 30);
{
fill(255, 0, 0);
rect(0, 160, newT*1.6, 30);
}
}
}
}
Will’s posted the PDF cheatsheet over here, and I’ll be starting a greenhouse project next week with a couple of Arduino sensors including this one, so I’d love to hear any advice from any hackers out there with a green thumb :) | http://antipastohw.blogspot.com/2010/05/ | CC-MAIN-2018-26 | refinedweb | 485 | 56.66 |
README
React Mini Chart ComponentsReact Mini Chart Components
A collection of light-weight mini chart components, built with SVG - with no dependencies.
InstallationInstallation
This package is pre-release and is yet to be published to NPM. To install, install from this repository.
npm install --save
SetupSetup
Note: The below is in ES2015, and can be transpiled with Babel or TypeScript.
You can import the whole library, or just the parts you want.
import ReactMiniChartComponents from 'react-mini-chart-components'; const Gauge = ReactMiniChartComponents.Gauge;
Or, import just the chart(s) you want.
import {Gauge} from 'react-mini-chart-components';
GaugeGauge
The Gauge component has the following configuration. All are optional.
Below is a
'half-gauge', with a value of
15, colored
'LimeGreen' and with a width of
'5em'.
<Gauge type='half-gauge' value={15}
ExampleExample
To see these components in use, you can view app.jsx. | https://www.skypack.dev/view/react-mini-chart-components | CC-MAIN-2021-43 | refinedweb | 145 | 51.04 |
Your browser does not seem to support JavaScript. As a result, your viewing experience will be diminished, and you have been placed in read-only mode.
Please download a browser that supports JavaScript, or enable it if it's disabled (i.e. NoScript).
Edit: Here is a solution by Maxime:
import c4d
import urllib2
import os
f = os.path.join(os.path.dirname(c4d.storage.GeGetStartupApplication()), "resource", "ssl", "cacert.pem")
urllib2.urlopen("", cafile=f)
This is a SSL certificates issue
Hi, @merkvilson thanks for contacting us.
Unfortunately, this is an already a known issue and it's not fixed yet.
You may find some workaround by googling the error message, but this indeed introduce some security issue so we can't recommend this as a correct workaround.
Cheers,
Maxime.
Does this apply only to Mac devices?
Ok. It seems like this problem occurs only on Mac devices.
Yes the SSL error is only a mac issue
It seems like there are two ways of solving this problem.
pip install --upgrade certifi
open /Applications/Python\ 3.6/Install\ Certificates.command
Is it possible to execute these codes directly from python plugin?
As I remember, there is a pip module which can be imported via import pip command in python code and then install the desired modules but will it work with C4D's python implementation?
Hi @merkvilson, I indeed overlooked the issue, since in all topic it's only Python 3.6, I didn't tried.
But in MacOs, certifi is not installed (which cause the issue), but you can directly get the certificate file to make it works as expected. Here the code which works on mac and windows.
import c4d
import os
import ssl
import urllib
f = os.path.join(os.path.dirname(c4d.storage.GeGetStartupApplication()), "resource", "ssl", "cacert.pem")
context = ssl.create_default_context(cafile=f)
urllib.urlopen("",context=context)
Or with urllib2
@m_adam
Thanks, Maxime! I'll check it as soon as one of the beta testers will be available online to test it out.
I'm getting this error message on windows pc. I guess this is expected behavior on windows, right?
urllib2.HTTPError: HTTP Error 405: METHOD NOT ALLOWED
urllib2.HTTPError: HTTP Error 405: METHOD NOT ALLOWED
@merkvilson said in C4D's python implementation on Mac and PC:
urllib2.HTTPError: HTTP Error 405: METHOD NOT ALLOWED
Does it happen on all websites? Or only a specific one? Could you try to open
Here it's working nicely on windows/mac
I tried again and it worked perfectly.
I'm not sure what caused the previous problem. I probably did something wrong.
I'm still testing it on my windows pc, and I'm not getting any errors.
I guess this thread will be marked as solved in a few minutes
Code worked on all of my beta testers' mac and windows devices.
Thanks, Maxime!
Hey everyone,
i am having this problem where I want to download an assets (zipfile) from a private repo on GitHub.
I've started to develop it with requests from there everything worked fine... now I am trying to port it to urllib2 for C4D but it doesn't work anymore...
requests
urllib2
'Accept': 'application/octet-stream' will always result in: HTTP Error 415: Unsupported Media Type
'Accept': 'application/octet-stream'
HTTP Error 415: Unsupported Media Type
If I get rid of 'Accept': 'application/octet-stream' it will give me the application/json
application/json
headers = {
'Authorization': 'token %s' % (MYTOKEN),
'Accept': 'application/octet-stream',
}
url = ''
request = urllib2.Request(url, headers=headers)
response = urllib2.urlopen(request)
print response.read()
#GIVES ME:
#urllib2.HTTPError: HTTP Error 415: Unsupported Media Type
Any idea how to avoid the HTTP Error 415 and download the zip to disk?
Really hard to find something about this anywhere...
HTTP Error 415
Any help is appreciated.
Thanks,
Lasse
Hi Lasse,
you're probably having some issues with authorization.
You're getting a JSON response which may indicate that your requested URL cannot be delivered.
What's the content of the reponse if you print it?
Does it happen with a public repo as well?
Best,
Robert
@mp5gosu
Hi Robert!
Haven't tested it with a public repo yet, but I'll bet its going to work ... the print of the response gives me:
print
response
urllib2.HTTPError: HTTP Error 415: Unsupported Media Type
which I can't find much info for ...
By the way, the URL you provided doesn't reflect GitHubs scheme. It is actually
@mp5gosu Yep, the link is coming from the provided JSON data. Also removing the repo from the link gives me the same error!
Is it possible that this has something to do with 'Accept': 'application/octet-stream'? Isn't it possible to stream with urllib2?
repo
Hey Lasse,
sorry - removing "repos" is actually wrong when using api access, my bad.
I can now also comprehend your problem. I'm going to dig a little deeper. | https://plugincafe.maxon.net/topic/11370/urllib2-urlopen-fails-on-c4d-for-mac/?lang=en-US&page=1 | CC-MAIN-2022-27 | refinedweb | 820 | 67.45 |
public class TFRecordIO extends java.lang.Object
PTransforms for reading and writing TensorFlow TFRecord files.
For reading files, use
read().
For simple cases of writing files, use
write(). For more complex cases (such as ability
to write windowed data or writing to multiple destinations) use
sink() in combination with
FileIO.write() or
FileIO.writeDynamic().
clone, equals, finalize, getClass, hashCode, notify, notifyAll, toString, wait, wait, wait
public static final Coder<byte[]> DEFAULT_BYTE_ARRAY_CODER
public static TFRecordIO.Read read()
PTransformthat reads from a TFRecord file (or multiple TFRecord files matching a pattern) and returns a
PCollectioncontaining the decoding of each of the records of the TFRecord file(s) as a byte array.
public static TFRecordIO.ReadFiles readFiles()
read(), but reads each file in a
PCollectionof
FileIO.ReadableFile, returned by
FileIO.readMatches().
public static TFRecordIO.Write write()
PTransformthat writes a
PCollectionto TFRecord file (or multiple TFRecord files matching a sharding pattern), with each element of the input collection encoded into its own record.
public static TFRecordIO.Sink sink()
FileIO.Sinkfor use with
FileIO.write()and
FileIO.writeDynamic(). | https://beam.apache.org/releases/javadoc/2.29.0/org/apache/beam/sdk/io/TFRecordIO.html | CC-MAIN-2021-43 | refinedweb | 173 | 53.07 |
17 February 2006 06:48 [Source: ICIS news]
SINGAPORE (ICIS news)--PTT Phenol, a Thai joint-venture, could build a bisphenol-A (BPA) unit downstream of its phenol project, a company source said.
The project at Mab Ta Phut, Rayong Province, ?xml:namespace>
It was unclear when the study would be completed.
PTT Phenol is building a 200,000 tonne/year phenol plant, which will start commercial production in mid-2008, at the same site.
The project will be based on UOP’s technology, and Foster Wheeler is the project management consultant. Benzene and propylene feedstocks will be sourced from the company’s shareholders Aromatics Thailand Co (ATC) and PTT Chemicals.
PTT Pcl, the kingdom's oil, gas and petrochemicals major, and PTT Chemicals each hold a 40% stake in PTT Phenol while ATC has a 20% | http://www.icis.com/Articles/2006/02/17/1042978/ptt-phenol-mulls-downstream-bpa-project.html | CC-MAIN-2015-18 | refinedweb | 137 | 70.13 |
This is ready for reading. You can test them with the macro FD_ISSET(), below.
Before progressing much further, I'll talk about how to manipulate these sets. Each set is of the type fd_set. The following macros operate on this type::
#include <sys/time.h>
#include <sys/types.h>
#include <unistd.h>
#define STDIN 0 /* file descriptor for standard input */
main()
{.
One final.
SELECT is an SQL command for retrieving data from a (relational) database. The simplest SELECT statement is this:
SELECT * FROM example
SELECT * FROM example
The statement above tells the database to return all data in the table called example, with the rows in no particular order. To limit the amount of data that is returned you can specify which fields in the table to return, and you can set conditions on the rows that are returned with a WHERE clause.
example
SELECT nickname, firstname, lastname FROM example WHERE nickname='nate'
SELECT nickname, firstname, lastname FROM example WHERE nickname='nate'
The statement above will return all rows that have the value 'nate' in the field nickname. Each returned row only contains data from the fields nickname, firstname and lastname.
'nate'
nickname
SELECT nickname, firstname, lastname FROM example ORDER BY lastname
SELECT nickname, firstname, lastname FROM example ORDER BY lastname
If you include the ORDER BY clause, then the rows will be sorted by the field(s) in the ORDER BY clause. If you add DESC after the field name(s), then the rows are sorted in descending order.
You can do much more with SELECT statements, such as selecting from more than one table (called joining), but unfortunately the syntax sometimes varies depending on which database server you use.
See LIKE for information on how to perform pattern matching.
Se*lect" (?), a. [L. selectus, p. p. of seligere to select; pref. se- aside + levere to gather. See Legend.]
Taken from a number by preferance; picked out as more valuable or exellent than others; of special value or exellence; nicely chosen; selected; choice.
A few select spirits had separated from the crowd, and formed a fit audience round a far greater teacher.
Macaulay.
© Webster 1913.
Se*lect", v. t. [imp. & p. p. Selected; p. pr. & vb. n. Selecting.]
To choose and take from a number; to take by preference from among others; to pick out; to cull; as, to select the best authors for perusal.
Milton.
The pious chief . . .
A hundred youths from all his train selects.
Dryden.
Log in or registerto write something here or to contact authors.
Need help? accounthelp@everything2.com | http://everything2.com/title/Select | CC-MAIN-2016-50 | refinedweb | 425 | 63.29 |
(IDEA newbie here, sorry if this is a FAQ or a silly question, but I didn't find a solution with a half an hour of googling)
I just updated to 10.0.2 and my Scala stuff stopped working because now IDEA will not recognize my Scala object as a runnable unless the first character of the file is uppercase.
Here's how I can reproduce the issue:
- Create a new project (from scratch) and assign it the Scala 2.8.1 facet. Scala distribution resides in scala/ in my home directory. I use OSX if that makes any difference.
- I create a new Scala class Test in the project src directory, and replace the autogenerated contents with the three lines:
object Test {
def main(args: Array[String]) = true
}
At this point the new class shows up in the project view accompanied with the yellow "O" icon, and the "Run Test.scala" option is not available in the popup menu of the file.
Now, if I capitalize the first letter of the word "object" above, ie. I have "Object Test ...", then the icon changes to the red Scala logo with a cog, and I can run it in the popup menu. However, the attempt will naturally fail with the error:
.../IdeaProjects/Test/src/Test.scala:1: error: value Test is not a member of object java.lang.Object
Object Test {
^
Process finished with exit code 1
If I change the uppercase O back to lowercase, the icon changes back to yellow "O" and I can no longer run the class/project. When I trigger the previously generated run configuration IDEA tells me "Error running Test.scala: Scala script file not found." even though I did not change the name of the file or the class.
Anyone know what's wrong?
(IDEA newbie here, sorry if this is a FAQ or a silly question, but I didn't find a solution with a half an hour of googling)
Alex,
The issue you're seeing have to do with the difference between Scala scripts (no package declaration and "loose" code at the top level like defs and vals or plain old evaluable expressions).
Scala scripts are signified by the Scala stairway icon (hard to make out) overlayed with the cogwheel. Only Scala scripts get the green Run arrow.
When you write
object Test { def main(args: Array[String]): Unit = { ... } }
... you're not writing a script. However, when you change that to
Object Test { def main(args: Array[String]): Unit = { ... } }
...you're changing the entire interpretation of this code. It is now (syntactically) deemed a script, but it will not compile (so technically it's not Scala at all, it's just a bunch of text...).
Randall Schulz
Thanks for the reply, that makes sense.
But I still don't understand how do I create the iconic hello world Scala object that is runnable in IDEA 10.0.2?
A file containing this:
-==-
println("Hello, world.")
-==-
That's a Scala script.
If you want to compile it to produce a runnable Scala program, use something like this:
-==-
package stone.alex.world
object Hello {
def
main(args: Array[String]): Unit =
println("Hello, world.")
}
-==-
Randall Schulz
Okay, I guess I asked the wrong question, sorry :)
How do I create a run configuration for the program? Ie. I want it to do what
scala stone.alex.world.Hello
does in the command line.
Again, apologies, these are trivial issues but I'm a total IDEA newbie and I can't seem get the run configuration dialog to do what I want.
Right mouse button menu -> "Run Hello.main()" entry.
Pressing Ctrl-Shift-F10 while you are in the body of the main() method should work as well.
I think the previous example that you quoted ("def main(args: Array[String]) = true") does not have the right signature for a Scala main() method, but I'm not 100% sure.
Note that there is a dedicated forum for the IntelliJ Scala plugin:
-tt
Thanks for the tip about the Scala plugin forum, I'll ask further questions there. | https://intellij-support.jetbrains.com/hc/en-us/community/posts/206256109-Weird-Scala-issue-with-10-0-2 | CC-MAIN-2020-34 | refinedweb | 681 | 80.41 |
j have. For instance:
<div data-</div>
There is a CSS selector we can use in jQuery to select that:
$("[data-whatever='elephant-eyeballs']");
There are variations on the attribute selector, like instead of
= you can do
^= to indicate “starts with this value”. But for some reason, CSS doesn’t have != or “not equal to this value”. jQuery does! That’s an example of a jQuery selector extension.
There are lots of these selector extensions. A few we specifically talk about in this screencast:
- :eq() – a 0-indexed version of :nth-child()
- :even – shortcut for :nth-child(even)
- :gt(n) – select elements with a greater index than n. Also a shortcut for a more complex :nth-child() forumla.
Possibly the most useful selector extension of all is :has() – which limits the selection to elements which contain what you pass this pseudo selector (or is it a pseudo pseudo selector :) It’s likely that someday in the future CSS will have something like this for us (I think it’s going to be like
pre ! code) but that’s a long way off. Unfortunately I don’t make a very compelling argument for it in this screencast, but you’ll know when you need it! For instance “Select all .module’s that contain an h3.sports-bar” – that kind of thing.
You can also make your own selection extensions if you wish. Here’s an article on that. The example is to make
:inview which selects an element only if it’s visible on the page at the current scroll position. It would be like this:
jQuery.extend(jQuery.expr[':'], { inview: function (el) { var $e = $(el), $w = $(window), top = $e.offset().top, height = $e.outerHeight(true), windowTop = $w.scrollTop(), windowScroll = windowTop - height, windowHeight = windowTop + height + $w.height(); return (top > windowScroll && top < windowHeight); } });
Select and do, select and do.
Select and do, select and do :)
Select and do, select and do… I couldn’t help myself ;)
Select and do, select and do, select and do
Thanks, Chris! I didn’t know we can make our own selector extensions in jQuery before this episode. | https://css-tricks.com/lodge/learn-jquery/06-jquery-selector-extensions/ | CC-MAIN-2022-05 | refinedweb | 351 | 66.44 |
Joe Mocker's Weblog How'd you know I was blogging at you if you weren't blogging at me? 2014-08-06T11:25:05+00:00 Apache Roller Running VirtualBox VMs as Services in Windows me 2010-03-04T15:07:00+00:00 2010-03-04T23:07:00+00:00 <p.<br /></p> <p> Well, I just found this handy little tool called <a href="">VBoxVmService<.</p> <p.</p> <p>The VMs start up at boot, and I can access them with Windows Remote Desktop Connection client. <br /></p> <p><br /></p> Stupid VirtualBox & ZFS Tricks me 2010-02-23T09:32:28+00:00 2010-02-23T17:32:28+00:00 <p>So the other day I had the idea to flip things around, instead of running a Windows 7 guest within an OpenSolaris host (with virtualbox) on my desktop, run an OpenSolaris guest within a Windows 7 host.</p> <p.</p> <p.<br /></p> <p>Now the interesting part(s), first hurdle, how do I get VirtualBox to recognize the physical disks the ZFS pool resides on. After a little web searching, it ended up being pretty easy, the key is to use the <tt>VBoxManage internalcommands createrawvmdk</tt> command that VirtualBox provides.</p> <p>Using this command, you can create a vmdk definition that points to a physical disk or partition. In my case, I wanted to point to the entire disk, so I used the following commands.</p> <pre>VBoxManage internalcommands createrawvmdk -filename C:\\Users\\Mocker\\.VirtualBox\\HardDisks\\PhysicalDrive2.vmdk" -rawdisk \\\\.\\PhysicalDrive2 -register VBoxManage internalcommands createrawvmdk -filename C:\\Users\\Mocker\\.VirtualBox\\HardDisks\\PhysicalDrive3.vmdk" -rawdisk \\\\.\\PhysicalDrive3 -register</pre> <p>Now, here's the rub, in order to do this, <tt>VBoxManage</tt> needs to be run as administrator. Windows 7 apparently has some pretty tight restrictions on who/what is allowed to access raw disks. One of the easier ways to do this is to just run a command shell as administrator, which will then execute <tt>VBoxManage</tt> as administrator.</p> <p>I did this by simply clicking the Start menu, entering "cmd" in the search box, which will present cmd.exe as a result, right click on cmd.exe and select "Run as Administrator".</p> <p. </p> <p>Note, same rub as above applies, I needed to run VirtualBox as administrator. This is getting to be a drag. </p> <p>Finally, the moment of truth, time to fire up the OpenSolaris guest, and see if it will recognize the ZFS pool...</p> <p>After the guest booted up, I logged in, using the <tt>format</tt> command, I could see the guest recognized the new disks, good.</p> <p> Next, I ran zpool import, and, drum roll, yes, it indeed found my ZFS pool! </p> <p>Finally, I ran <tt>zpool import <poolname></tt> and voila! OpenSolaris happily imported the pool - lock, stock and barrel --</p> <pre># zpool status storage pool: storage state: ONLINE scrub: none requested config: NAME STATE READ WRITE CKSUM storage ONLINE 0 0 0 mirror-0 ONLINE 0 0 0 c8d1s0 ONLINE 0 0 0 c7d1s0 ONLINE 0 0 0</pre> <p.</p> <p>I managed to fix this pretty easily though, I simply booted back into OpenSolaris on bare metal (yeah, I actually created a dual boot system), then though a process of <tt>zfs detach</tt>, <tt>format</tt>, <tt>zfs attach</tt>, I relabeled the disks and Windows 7 was much happier.<br /></p> <p><br /></p> How to Shrink a mirrored ZFS rpool me 2010-01-29T14:25:59+00:00 2010-01-29T22:25:59+00:00 <p>The other day I wanted to <i>shrink</i> the size of my OpenSolaris ZFS root pool.The root pool is generally pretty small, and when possible I try to keep real data out of the root pool. In the case of target system, it has 4 disks and I have separate mirrored data pool. So I've got nearly 1 TB on two disks (root pool mirroring) just sitting idle, which could be used for something like say installing other O/Ses on bare iron (yeah, sometimes I still like to do that).<br /></p> <p>Anyways, maybe I am pathetic at web searches but after a few minutes of searching I could not find an answer. So I thought about it for a while and came up with a plausible way to do it, without the need of a live CD or usb stick or anything (did I mention I'm lazy, and didn't really feel like burning a CD)</p> <p>The method I came up with is this, I'll follow up a summary with more detailed steps. Note that although this procedure is to shrink a mirrored root pool, you can probably use the same method on a non-mirrored pool, as long as you have a spare hard drive or partition somewhere.<br /></p> <ol> <li>Break the root pool mirror</li> <li>Create a temporary root pool, resized appropriately, and boot to it</li> <li>Destroy the real root pool, resize it, and boot back to it</li> <li>Destroy the temporary root pool.</li> <li>Reattach the mirror to the resized root pool</li> </ol> <p>It sounds pretty simple, but there are a lot of steps involved. So here are the details.</p> <p>In my case, I have a mirrored root pool made of the two devices</p> <blockquote> <p>c2t0d0s0<br />c2t1d0s0</p> </blockquote> <p> </p> <p> <b>Break the root pool mirror.</b> <br /></p> <blockquote> <pre>zpool detach rpool c2t1d0s0 </pre> </blockquote> <b>Create a temporary root pool, resized appropriately, and boot to it.</b> <p>First, you need to resize the Solaris fdisk partition, I usually just use the fdisk option in format, the general outline for this procedure is the following:</p> <ol> <li>Run format</li> <li>Select the disk you just detached (c2t>Now, create a temporary root pool, I'll call it tpool<br /> <blockquote> <pre><dev>zpool create -f tpool c2t1d0s0</dev></pre> </blockquote> <p>Copy the data from the root pool to the temporary pool with ZFS send & receive<br /><dev><dev /></dev></p> <blockquote> <pre><dev><dev>zfs snapshot -r rpool@shrink zfs send -vR rpool@shrink | zfs receive -vfd tpool </dev></dev></pre> </blockquote> <p>Now, before you can boot to the temporary pool, there's a couple setting you need to change so its identified as tpool properly. <dev><dev /></dev></p> <p><dev><dev>First, change the boot sign of the temporary pool </dev></dev></p> <blockquote> <pre><dev><dev>rm /tpool/boot/grub/bootsign/pool_rpool touch /tpool/boot/grub/bootsign/pool_tpool </dev></dev></pre> </blockquote> <p><dev><dev>Now, set the bootfs option on the pool </dev></dev></p> <blockquote> <pre><dev><dev>zfs set bootfs=tpool/ROOT/opensolaris-5 tpool </dev></dev></pre> </blockquote> <p><dev><dev>And finally, make sure to install GRUB on the disk </dev></dev></p> <blockquote> <pre><dev><dev>cd /boot/GRUB /sbin/installgrub stage1 stage2 /dev/rdsk/c2t1d0s0<dev> </dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>Now, reboot. Make sure do do a full PROM boot instead of the newer Quick boot </dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>reboot -p </dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>At grub, edit the boot entry - by typing 'e' - for the BE you want to boot into, change references from rpool to tpool. In my case the boot commands ended up looking like</dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>findroot (pool_tpool,0,a) bootfs tpool/ROOT/opensolaris-5 kernel$ /platform/i86pc/kernel/$ISADIR/unix -B $ZFS-BOOTFS,console=text module$ /platform/i86pc/$ISADIR/boot_archive</dev></dev></dev></pre> <pre><dev><dev><dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>Then type 'b' to boot. When the system has completed booting, you should be booted off the temporary pool. You can verify this by</dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>df -h / </dev></dev></dev></pre> </blockquote> <p><b>Destroy the real root pool, resize it, and boot back to it</b>. <br /></p> <p>First off, before you can destroy the root pool, the system might have some active references into the pool, like for swap and dump,<dev><dev><dev> and maybe some datasets like /export, you need to deactivate these before destroying the rpool. </dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>swap -d /dev/zvol/dsk/rpool/swap dumpadm -d /dev/zvol/dsk/tpool/dump zfs set mountpoint=none rpool/export zfs set mountpoint=none rpool/export/home </dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>Now you should be able to destroy the original rpool </dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>zpool destroy rpool </dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>If you get "busy" errors, try destroying the datasets which should identify why its busy, and deactivate them. </dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>zfs destroy -r rpool zpool destroy rpool </dev></dev></dev></pre> </blockquote> <p><dev><dev><dev>With the original root pool destroyed, you can now repartition the device in the original pool to the new size.</dev></dev></dev></p> <p> </p> <ol> <li>Run format</li> <li>Select the disk you just detached (c2t> <p><dev><dev><dev>With the Solaris partition resized, recreate rpool </dev></dev></dev></p> <blockquote> <pre><dev><dev><dev>zfs create rpool c2t2d0s0 </dev></dev></dev></pre> </blockquote> <p>And reverse the process of migrating to the temporary pool by moving the data back to the new root pool <br /></p> <blockquote> <pre><dev><dev><dev><dev>zfs send -vR tpool@shrink | zfs receive -vfd rpool </dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev><dev>Once again, set the bootfs option on the pool </dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev>zfs set bootfs=rpool/ROOT/opensolaris-5 rpool </dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev><dev>Reinstall GRUB</dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev>cd /boot/grub /sbin/installgrub stage1 stage2 /dev/rdsk/c2t2d0s0<dev> </dev></dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev><dev><dev>And do a full PROM boot<br /> </dev></dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev><dev>reboot -p </dev></dev></dev></dev></dev></pre> </blockquote> <p><b>Destroy the temporary root pool & remirror root pool.</b><br /></p> <p><dev><dev><dev><dev><dev>At this point the rpool is resized smaller, but not mirrored. To mirror again destroy the temporary pool </dev></dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev><dev>zpool destroy tpool </dev></dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev><dev><dev>Attach tpool device to the rpool </dev></dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev><dev>zpool attach rpool c2t2d0s0 c2t1d0s0</dev></dev></dev></dev></dev></pre> </blockquote> <p><dev><dev><dev><dev><dev><dev> <tdev> And finally install grub once again for good measure </tdev></dev></dev></dev></dev></dev></dev></p> <blockquote> <pre><dev><dev><dev><dev><dev><dev><tdev>/sbin/installgrub stage1 stage2 /dev/rdsk/c2t1d0s0</tdev></dev></dev></dev></dev></dev></dev> </pre> </blockquote> <pre><p>Voila. </p></pre> Paranoia and Java Cryptography me 2009-05-16T08:06:05+00:00 2009-05-16T15:06:05+00:00 <p> I started looking at rewriting a web app I wrote a while ago that does encryption in Java. This time trying to pay more attention to crossing the T's and dotting the I's, and I'm realizing encryption in Java is tricky. <p> <p> The JCE itself makes things pretty easy, however where it gets a little tricky is how to deal with discarded information - passphrases, private keys, unencrypted data - when you are done with it. Unlike C code, for example, where you can simply zero out portions of memory when you are done with it, Java has very little available to do the same thing. <p> <p> And it can get worse, when you are dealing with Strings, for example, they can get stuffed into a master String table - via String.intern(), squirreled away in the far reaches of the JDK, with no ability to destroy them. Thus, at the very least, never convert any sensitive information to a String, if you want any hope of ever clearing it out of memory. <p> <p> The JCE designers seem to have been pretty keen to this, and provide interfaces that never use Strings as any parameters that could be sensitive. A good foundation, however dealing sensitive data elsewhere can be tricky. <p> <p> A simple example, back to the web app, is dealing with parameters in a servlet. Say you have a servlet which takes a passphrase and a chunk of text and encodes it. Normally you retrieve parameters in a servlet with <tt>HttpServletRequest.getParameter(String name)</tt>. The problem: <tt>getParameter()</tt> returns a String, and thus could get stuffed into the JVMs String table for ever more. <p> <p> Although it would be unlikely for someone to gain access to the String table, and then figure out which String actually represented a passphrase, the paranoid side in me makes me a little nervous to allow that sensitive information to exist out of my control. <p> <p> Even worse, the unencrypted data that you want to encrypt. Again, if you retrieve it with <tt>getParameter()</tt>, you won't be able to fully discard the unencrypted data until you restart the JVM. </p> <p> I haven't quite figured out a plan for how paranoid I want to be. One thought would be to instantiate a new ClassLoader to manipulate sensitive data. Presumably when you get rid of the ClassLoader all the classes (including the String table) would at least be eligible for garbage collection. </p> My Own Private Cloud-aho - The Details me 2009-04-16T13:54:17+00:00 2009-04-16T20:54:17+00:00 <P STYLE="margin-bottom: 0in"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>Network Layout</U></FONT></P> <P STYLE="margin-bottom: 0in"> <a href="/mock/resource/cloud/CloudInfrastructure.png"><img class="imgr" src="/mock/resource/cloud/CloudInfrastructure-sm.png"/></a> <FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>Install Servers</U></FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>DHCP Servers<BR></U></FONT> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">The one other interesting thing to note is that the pair of DHCP servers are configured as a <I>cluster</I>. <FONT FACE="monospace">SUNWfiles</FONT> datasource. Neither <FONT FACE="monospace">SUNWbinfiles</FONT> nor <FONT FACE="monospace">SUNWnisplus</FONT> will work. </FONT> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">A couple of other things to be aware of when <I>clustering</I> DHCP is to make sure you set the <FONT FACE="monospace">OWNER_IP</FONT> parameter in each daemon's local <FONT FACE="monospace">dhcpsvc.conf</FONT> file to a comma separated list of all the IP addresses of all the interfaces that will serve DHCP requests on all servers. Also make sure you set the <FONT FACE="monospace">RESCAN_INTERVAL</FONT> to a reasonable value for you, in our case we just set it to 1 minute. Both of these values can be updated with the <FONT FACE="monospace">dhcpconfig -P</FONT> command.</FONT></P> <PRE STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> <FONT FACE="monospace"><FONT SIZE=2>dhcpconfig -P OWNER_IP=192.168.78.14,192.168.78.16,192.168.76.25,192.168.76.25 </FONT></FONT> <FONT FACE="sans"><FONT SIZE=2><FONT FACE="monospace">dhcpconfig -P RESCAN_INTERVAL=1</FONT></FONT> </FONT> </PRE> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>The Hypervisors</U></FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">Not much in the way of customizations to Dom-0 or the Hypervisor. Obviously in Dom-0 we disable as many unnecessary services as possible. Since these are servers, we shut down all of the Desktop related services like the X server and the like. </FONT> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">We also limit Dom-0 memory to only 2G using the <FONT FACE="monospace">dom0_mem</FONT> parameter to the hypervisor. This might be a little aggressive since ZFS is memory hungry, but we want to try to keep as much memory available for the guest domains as possible, and we haven't seen a problem with this yet.</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">We also set the hypervisor console to com1, in case we need to break into the console for any sort of debugging (knock on wood we don't have to do that.)</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">Both these parameters are set from the GRUB boot commands</FONT></P> <P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> <FONT FACE="monospace"><FONT SIZE=2>kernel$ /boot/$ISADIR/xen.gz com1=9600,8n1 console=com1 dom0_mem=2G</FONT></FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">We also use the vanity device naming capabilities of Solaris - via <FONT FACE="monospace">dladm</FONT> -.</FONT></P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"><FONT FACE="monospace">dladm rename-link e1000g0 fe0</FONT> <FONT FACE="monospace">dladm rename-link e1000g1 be0</FONT></PRE><P STYLE="margin-bottom: 0in; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">Finally, while we're talking about Live Migration, its something we need to enable in <FONT FACE="monospace">xend</FONT>. A couple of simple SMF changes handle that.</FONT></P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"> <FONT FACE="monospace"><FONT SIZE=2>svccfg -s xend setprop config/xend-relocation-address = 192.168.78.21</FONT></FONT> <FONT FACE="monospace"><FONT SIZE=2>svccfg -s xend setprop config/xend-relocation-hosts-allow = astring: \\</FONT></FONT> <FONT FACE="monospace"><FONT SIZE=2> \\"\^localhost$\^192\\.168\\.78\\.[0-9]\*$\\"</FONT></FONT> </PRE> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">For what its worth, we currently have 22 Hypervisors in the cloud, clearly not yet a huge deployment, yet.</FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>Unified Storage Cluster</U></FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">As far as how we are using them, well, they are used in a few capacities. First, as I mentioned earlier, they are used to house the DHCP servers' shared datastore. We also use them to house various administrative tools and bits.</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">This is another critical piece that makes some of the cool features of xVM Xen, like Live Migration, possible. Live Migration is the process of moving a running virtual server from one physical host to another. Did you read that, a <B>running virtual server</B><SPAN STYLE="font-weight: medium">!.</SPAN></FONT></P> <P STYLE="margin-bottom: 0in; font-weight: medium; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in; font-weight: medium; text-decoration: none"> <FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; font-weight: medium; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>Putting it All Together</U></FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">That's a summary of all the pieces, now how does it all fit together. Here's a diagram that shows all the interactions.</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"> <a href="/mock/resource/cloud/CloudInstantiation.png"><img class="imgr" src="/mock/resource/cloud/CloudInstantiation-sm.png"/></a> <FONT FACE="sans".)</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> <FONT FACE="sans"><FONT SIZE=2><I>These commands will create a new project called <FONT FACE="monospace">appl-1,</FONT> and then clone a master image to the project as <FONT FACE="monospace">vm-1</FONT>.</I></FONT></FONT></P> <P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> </P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"><FONT FACE="monospace">domu-project appl-1</FONT> <FONT FACE="monospace">domu-clone masters/osol-0811@version-01 appl-1/vm-1</FONT></PRE><P STYLE="margin-bottom: 0in; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">From here, the usual Xen commands are used to define a domain, that is <FONT FACE="monospace">xm create</FONT> or <FONT FACE="monospace">virsh create</FONT> commands. As part of the domain configuration, we specify the attached iSCSI LUN for the virtual disk for the domain. Again, this is simplified though scripting by using a pre-defined XML template for domain creation.</FONT></P> <P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> <FONT FACE="sans"><FONT SIZE=2><I>This example shows the creation of a paravirtualized guest - <FONT FACE="monospace">pv</FONT>, which has both a Public and Backend interface - <FONT FACE="monospace">fe-be</FONT>, on the Odd network segment - <FONT FACE="monospace">odd</FONT>, with 1G of memory - <FONT FACE="monospace">1024</FONT>, and 1 CPU - <FONT FACE="monospace">1</FONT>.</I></FONT></FONT> </P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"><FONT FACE="monospace">domu-init pv fe-be odd 1024 1 appl-1/vm-1</FONT></PRE><P STYLE="margin-bottom: 0in; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans".</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans" <I>cluster</I>, since so much relies on it. Once again, this is scripted.</FONT></P> <P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> <FONT FACE="sans">This example shows the assignment of a Public and Backend interface to the new guest</FONT> </P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"> <FONT FACE="monospace">domu-assign-ip appl-1:vm-1 fe0 vm-host1 192.168.76.71</FONT> <FONT FACE="monospace">domu-assign-ip appl-2:vm-1 be0 vm-host1-be 192.168.78.71</FONT></PRE><P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">And that's it. At this point, we're ready to fire up the guest.</FONT></P> <PRE STYLE="margin-left: 0.49in; text-decoration: none"><FONT FACE="monospace">virsh start appl-1:vm-1</FONT> <FONT FACE="monospace">virsh console appl-1:vm-1</FONT></PRE><P STYLE="margin-left: 0.49in; margin-bottom: 0in; text-decoration: none"> </P> <P STYLE="margin-bottom: 0in"><FONT FACE="sans"><U>Wrap Up</U></FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">That's about it for the basic details of what we've done. Its all been working incredibly well, especially considering we're running about ¾ development code everywhere.</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">Before I forget, I would really like to thank the xVM Xen team for their support while we've been setting things up and tinkering around, they have all been vary helpful and responsive to my questions on <A HREF="mailto:xen-discuss@opensolaris.org">xen-discuss@opensolaris.org</A> as well as private threads. Mark Johnson deserves a special mention since I glommed onto him the most.</FONT></P> <P STYLE="margin-bottom: 0in; text-decoration: none"><FONT FACE="sans">Up next, a summary of how well we're doing on the previously outlined goals.</FONT></P> My Own Private Cloud-aho - The Goals me 2009-04-14T17:06:13+00:00 2009-04-16T20:54:32+00:00 <P STYLE="margin-bottom: 0in"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>For quite some time I have been informally tracking the progress of the xVM Xen team, playing with code drops on perhaps a quarterly basis, and doing benchmarking to determine if things are stable enough and efficient enough to run in a production data-center. </FONT></FONT> </P> <P STYLE="margin-bottom: 0in"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Probably about 6 months ago, performance looked pretty decent, and I implemented a tiny proof-of-concept “cloud” using some friends as guinea pigs. Its been running great for us. In fact, I just checked my virtual server and its been running for 245 days so far. Pretty stable indeed.</FONT></FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Recently I have had the opportunity to build a small cloud in one of our data-centers, expanding on the concepts from the POC into a more full-blown and mature solution, suitable for a near-production environment. Its hard to really say its production quality when so many moving parts are still pre-release.</FONT></FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>I'd like to share what I have done with xVM Xen, in hopes that it may be useful for others who wish do do the same. But before I get into the details, let me describe some of the goals that we hope to achieve with this new virtual environment.</FONT></FONT></P> <P STYLE="margin-bottom: 0in"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3><U>The Goals</U></FONT></FONT></P> <UL> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Replace the current virtual hosting environment<BR><BR>I hate to knock Solaris Containers since its a Sun technology, but we've struggled somewhat in our usage. The problem isn't really with the technology, but rather a scheduling problem when it comes to planned and unplanned maintenance. In our environment we typically load up 4-8 application zones within a single physical host. These applications are generally maintained by separate teams. When we want to do something like patch the system, the operations team needs to schedule downtime with all the application teams for the same time so patching can occur, and we've found that the operations team can spin quite a few cycles lining all the ducks in a row.<BR><BR>The model with xVM Xen is quite a bit different. Since each domain, or virtual server, is more or less independent of the others, the operations team need only schedule down time with a single team at a time. So although they are doing more scheduling and patching, overall we hope that we can reduce the amount of real time they are doing entirely.<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Simple, Rapid (Virtual) Server Creation<BR><BR>The operations team has things dialed in pretty well with JASS, Jumpstart, and so on, for installing physical hosts, but the creation of virtual servers isn't as streamlined. The hope is that though the use of newer technologies available with ZFS – snapshots and cloning for one – as well as with a network storage infrastructure -and iSCSI, we can really streamline the process so that we can spin up virtual servers within minutes. The goal is to make the gating factor be how fast the system administrator can type.<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Faster & Error Free Deployments<BR><BR>Although most of the applications behind sun.com don't require huge clusters of servers, we do have a few that span perhaps a dozen or more physical hosts. The problem is that typically the process of deploying a new version the application requires the same set of repeated steps for each instance of the application, introducing the fat-finger problem. What if, using those same ZFS and iSCSI technologies, we can install once, then clone the application for the other instances. As long as that initial install is done correctly, it can greatly reduce the possibility of errors when replicating the changes across the cluster.<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Easier Horizontal Expansion<BR><BR>Occasionally, during launches or even worse, DOS attacks, applications can get hit their capacity which result in reduced quality of service for everyone. It those cases, it would be great if we could instantly increase our capacity. Are there technologies that we could employ to do this easily? We think there are.<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Painless Migration to Newer, Bigger, Faster Hardware<BR><BR>Although we've tried to employ some best practices which attempt to separate the O/S and the application on different areas of the filesystem(s), it still isn't the easiest exercise to upgrade an application to a new chunk of hardware. Essentially it becomes another case where the application team has to spend some cycles installing the service on the new hardware, test, verify, yadda, yadda, yadda.<BR><BR>We think that the live migration capabilities of Xen have great potential here. Since the application would be installed in a virtual server, the process of upgrading simply becomes a push of the running application from one physical host to another. And, this could even be something the operations team does all by itself, unbeknownst to the application team at all!<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>Better Hardware Utilization<BR><BR>A while ago I gave a talk with the xVM Xen team about what we had done. I don't think I really explained this one correctly, because their initial comment was something about Solaris Zones providing the most efficient method of squeezing every ounce out of a physical host.<BR><BR>That's not really what this is about. Many of the physical hosts are really incredibly under utilized, perhaps peaking out at somewhere near 30% of sustained CPU used even at the 95<SUP>th</SUP> percentile. With hundreds of hosts running that way, we're really just wasting power, cooling and space, when we don't need to.<BR><BR>We're hoping that with the virtualization capabilities with xVM Xen provides, we can make the practice of doubling or tripling up applications on a physical host more common, increasing the sustained performance closer to somewhere between 60% to 80% and lowering our datacenter footprint overall. Where an application begins to run hotter, monitoring would help us decide to move it, via Xen live migration, to a less used, larger, or private physical host.<BR></FONT></FONT></FONT><BR> </P> <LI><P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>More Resiliency<BR><BR>What we are looking for here is for ways to be able to recover better from catastrophic failures. The server is on fire, how do we get the application off of it and up and running quickly on another physical host? How do we reduce the need for someone hands on to physically fix a problem on a piece of hardware. Again, we're hoping virtualization and other technologies will be helpful here.</FONT></FONT></FONT></P> </UL> <P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <BR> </P> <P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>I probably forgot a few goals but in general these are the bigger problems which we hope to solve with a more virtualized data-center. </FONT></FONT></FONT> </P> <P STYLE="margin-bottom: 0in; font-style: normal; font-weight: medium; text-decoration: none"> <FONT COLOR="#000000"><FONT FACE="Liberation Sans, sans-serif"><FONT SIZE=3>In my next post, I'll describe the infrastructure we built in detail. </FONT></FONT></FONT> </P> Changing WS 7 Admin Server Certificate me 2008-10-02T15:08:25+00:00 2008-10-02T22:08:25+00:00 <p> This is probably not supported and the WS team might slap me, but... Occasionally I install an instance of WS 7 on a host, and later rename the host. Usually because the server gets pushed to our production data-center, and thus, gets a new name. </p> <p>. </p> <p> Its a minor annoyance, but it bugs me none-the-less. So I poked around and figured out how to change it. Its pretty simple really, just use the <tt>certutil</tt> command-line tool included with WS 7. Specifically, here's what I did. </p> <p> Go to the Admin Server's config folder <p> <pre> cd $WS_HOME/admin-server/config </pre> <p> Add the Admin Server's <tt>bin</tt> folder to your path </p> <pre> setenv PATH=$WS_HOME/bin:$PATH </pre> <p> Delete the old certificate, named <tt>Admin-Server-Cert</tt> </p> <pre> certutil -L -d . -n Admin-Server-Cert </pre> <p> Create the new certificate, specifying the new hostname in the <tt>-s</tt> parameter </p> <pre> certutil -S -d . -n Admin-Server-Cert -t u,u,u -s "CN=some.host.com" -c Admin-CA-Cert -v 120 </pre> <p> Last, you will probably need to change any references to the old host-name to the new name in <tt>server.xml</tt> </p> Running VNC over SSH on Windows me 2008-09-09T10:54:09+00:00 2008-09-09T17:54:09+00:00 <p> Its been a long time coming but, in keeping with my other write ups on how to run VNC over SSH on <a href="">Solaris</a> and <a href="">OS X</a>, I've finally figured out a somewhat reasonable method on Windows. </p> <p> I had been using <a href="">STunnel</a>. </p> <p> So I worked out a method similar to on Solaris and OS X with SSH Port forwarding. Only in this case, using the <tt>plink.exe</tt> utility that comes with <a href="">PuTTY</a>. First, the script. For someone who actually knows VBScript and Windows Script Host, this is probably pretty trivial, but for me, one who's tried to stay clear of Windows development, it took a bit of hacking: </p> <pre> Dim WshShell Set WshShell = CreateObject("WScript.Shell")TightVNC</a>. It'll probably work with other clients, but you'd have to give that a try. </p> <p> So, how do you use it. As with the method on the other Operating Systems, SSH Public Key Authentication is used. On Windows, with PuTTY, this means firing up <tt>pagent.exe</tt> and loading your keys into the agent. I'll leave specifics to the reader, I think I've mentioned it in other entries though. </p> <p> After that, simply double click on the script, and a prompt will show up asking for the <tt>host:port</tt> of the VNC server to connect to. Give it a couple of seconds to make the necessary connections and boom, you should be presented with a dialog to enter the password of the VNC server. </p> <p> If anyone finds this interesting, and would like to add some enhancements, one thing I would like to have is a connection history. So I don't have to type as much. </p> Hacking Lightning me 2008-09-08T12:37:00+00:00 2008-09-08T19:37:00+00:00 <p> I've been using the <a href="">Lightning</a> calendaring extension for <a href="">Mozilla Thunderbird</a> for the last couple of months for basic calendaring and it works pretty awesome. </p> <p> After <a href="">Rama</a> hassled me about keeping my task list on my whiteboard at work, I decided to try out the Task lists functionality in Lightning. Its pretty light weight, easy to add a task, almost exactly what I need. </p> <p> However, one thing that bugged was the predefined views. There's about four or five including a "Not Started" view and a "Completed" view, but none for what I really wanted, a "Not Completed" view. So I decided to add one. </p> <p> What I really wanted was to add another view, but even though I appeared to tweak all the files necessary, it didn't work. So in the end I just changed the "Overdue" view to "Overview & Open". And it was really easy one I found my way around. </p> <p> The tweak involves modifying files in two jars; <tt>calendar.jar</tt> & <tt>calendar-en-US.jar</tt> both located in <tt>~/.thunderbird/(profile)/extensions/(uuid)/chrome</tt>. </p> <p> In <tt>calendar-en-US.jar</tt> I wanted to change the label of the button from "Overdue" to "Overview & Open". This is done by unzipping the file, editing <tt>locale/en-US/calendar/calendar.dtd</tt>. and rezipping the file, changing the line </p> <pre> <!ENTITY calendar.task.filter.overdue.label "Overdue"> </pre> <p> to </p> <pre> <!ENTITY calendar.task.filter.overdue.label "Overdue & Open"> </pre> <p> Now, in the second jar <tt>calendar.jar</tt>, I needed to change the logic for deciding what was "overdue". This logic is in <tt>content/calendar/calendar-task-view.js</tt> The original logic is in the lines </p> <pre> overdue: function filterOverdue(item) { // in case the item has no due date // it can't be overdue by definition if (item.dueDate == null) { return false; } return (percentCompleted(item) < 100) && !(item.dueDate.compare(now()) > 0); }, </pre> <p> Which you can see they make a specific provision for excluding things with no dueDate. Many of my tasks do not have a dueDate and so I want to see those too. So I changed the logic to </p> <pre> overdue: function filterOverdue(item) { return (percentCompleted(item) < 100 && (item.dueDate == null || !(item.dueDate.compare(now()) > 0))); }, </pre> <p> Once I repacked the jars, voila, I now have the behavior I desired. </p> Converting Oracle Dates to UNIX Epoch Dates me 2008-09-02T16:19:14+00:00 2008-09-02T23:19:14+00:00 <p> Cleaning off my whiteboard, and I want to write these down somewhere before I erase them... </p> <p> To convert from a UNIX date to Oracle date </p> <pre> TO_DATE('1970-01-01', 'YYYY-MM-DD') + UNIX_date_in_millis / 86400000 = Oracle_date </pre> <p> And to convert the other way </p> <pre> Oracle_date - TO_DATE('1970-01-01', 'YYYY-MM-DD') \* 86400000 = UNIX_date_in_millis </pre> <p> I don't remember why I needed this so long ago. </p> My Partition Resizing Exercise mock 2008-05-06T12:29:50+00:00 2008-05-06T19:29:50+00:00 <p> The other day I needed to install a non-Solaris operating system onto my Ultra 40 M2. But I still wanted Solaris and/or Nevada as my primary operating system. Even though there are a bunch of virtualization technologies out there including the newly aquired VirtualBox, I wanted to run the operating system on the "iron" as they say. </p> <p> The only problem, I had allocated 100% of all my disks to Solaris. I needed to decrease the fdisk partition on a couple of disks to make space to install other operating systems on. My disk layout was roughly </p> <pre> c1t0d0s0 - boot environment #1 root (ufs) c1t0d0s1 - boot environment #1 swap c1t1d0s0 - boot environment #2 root (ufs) c1t1d0s1 - boot environment #2 swap </pre> <p> And then a ZFS pool </p> <pre> NAME STATE READ WRITE CKSUM storage ONLINE 0 0 0 mirror ONLINE 0 0 0 c1t2d0 ONLINE 0 0 0 c1t3d0 ONLINE 0 0 0 mirror ONLINE 0 0 0 c1t0d0s7 ONLINE 0 0 0 c1t1d0s7 ONLINE 0 0 0 </pre> <p> My thought was that since <tt>c1t0d0</tt> and <tt>c1t1d0</tt> were only partially allocated to the zpool (and the boot disks), that I would target those for repartitioning. Which meant, first step was to rebuild the zpool without them. I didn't really need them for more space in the zpool anyways. </p> <p> First step, to break the mirror, on the partial space. This would allow me to make a "backup" of the pool while I rebuild it. </p> <pre> zfs detach c1t1d0s7 zpool create dataz c1t1d0s7 zfs snapshot -r storage@migration zfs send -R storage@migration > /dataz/storage.zfs </pre> <p> Now, destroy and recreate the main pool without the partial devices, and restore the data </p> <pre> zpool destroy storage zpool create storage mirror c1t2d0 c1t3d0 zfs receive -Fd storage < /dataz/storage.zfs </pre> <p> Easy enough. My pool no longer has the partial disks </p> <pre> mirror ONLINE 0 0 0 c1t2d0 ONLINE 0 0 0 c1t3d0 ONLINE 0 0 0 </pre> <p> Now onto resizing the target disks. The goal was to do this without doing a reinstall at all. I have been using Live Upgrade for a while to continually keep my system up to date with Nevada/SXDE releases, so I figured I could use Live Upgrade while resizing the fdisk partitions to accomplish this. </p> <pre> Boot Environment Is Active Active Can Copy Name Complete Now On Reboot Delete Status -------------------------- -------- ------ --------- ------ ---------- sol-nv-87 yes yes yes no - sol-nv-85 yes no no yes - </pre> <p> So, first disk. First step remove the LU Boot Environment </p> <pre> ludelete sol-nv-85 </pre> <p> Now, since nothing is using the disk, <tt>format</tt> and <tt>fdisk</tt> aren't going to complain. So, first thing, use <tt>fdisk</tt> to reduce the size of the Solaris partition. I just reduced it to 90% of the disk <p> <pre> fdisk c1t1d0p0 Total disk size is 36472 cylinders Cylinder size is 16065 (512 byte) blocks Cylinders Partition Status Type Start End Length % ========= ====== ============ ===== === ====== === 1 Solaris 1 32824 32824 90 </pre> <p> And then recreated the Solaris layout with <tt>format</tt> </p> <pre> Current partition table (original): Total disk cylinders available: 32822 + 2 (reserved cylinders) Part Tag Flag Cylinders Size Blocks 0 root wm 132 - 1437 10.00GB (1306/0/0) 20980890 1 swap wu 1 - 131 1.00GB (131/0/0) 2104515 2 backup wu 0 - 32821 251.43GB (32822/0/0) 527285430 3 unassigned wu 0 0 (0/0/0) 0 4 unassigned wu 0 0 (0/0/0) 0 5 unassigned wu 0 0 (0/0/0) 0 6 unassigned wu 0 0 (0/0/0) 0 7 home wm 1438 - 32821 240.41GB (31384/0/0) 504183960 8 boot wu 0 - 0 7.84MB (1/0/0) 16065 9 unassigned wu 0 0 (0/0/0) 0 </pre> <p> And finally, recreate the Boot Environment. I just made the new Boot Environment a copy of the current Boot Envionment, which happens to be Nevada/SXDE Build 87 <p> <pre> lucreate -m /:/dev/dsk/c1t1d0s0:ufs -m -:/dev/dsk/c1t1d0s1:swap -n sol-nv-87-copy </pre> <p> And activate it and boot into it </p> <pre> luactivate sol-nv-87-copy init 6 </pre> <p> All pretty straight forward so far. Now here comes what ends up being the sticky part. You would think that since I was booted into the new Boot Environment, that I could delete the old Boot Environment, and do to the same thing, but... </p> <pre> ludelete sol-nv-87 The boot environment <sol-nv-87> contains the GRUB menu. Attempting to relocate the GRUB menu. ERROR: No suitable candidate slice for GRUB menu on boot disk: </dev/rdsk/c1t0d0p0> INFORMATION: You will need to create a new Live Upgrade boot environment on the boot disk to find a new candidate for the GRUB menu. ERROR: Cannot relocate the GRUB menu in boot environment <sol-nv-87>. ERROR: Cannot delete boot environment <sol-nv-87>. Unable to delete boot environment. </pre> <p> That's what I was afraid of. The GRUB boot menu is sitting on the disk still. I searched around but could not find any way to move it. I found <a href="">Slava Leanovich's blog entry</a> which included instructions for moving the GRUB menu, and even though I tried that, <tt>ludelete</tt> still complained. </p> <p> Well, gulp, here goes nothing. I decided to try to trick Solaris and not tell it I messed with the Boot Environment. First I made a <tt>ufsdump</tt> of root on that disk </p> <pre> ufsdump 0f /extra/holding/root.dump /dev/dsk/c1t0d0s0 </pre> <p> Now with everything backed up from the disk, a big leap of faith. I went into <tt>fdisk</tt> and resized the Solaris parition as just before. </p> <p> And then into <tt>format</tt> to recreate the disk layout. Of course this time I get the warnings </p> <pre> format c1t0d0 selecting c1t0d0 [disk formatted] /dev/dsk/c1t0d0s0 is in use for live upgrade /. Please see ludelete(1M). /dev/dsk/c1t0d0s1 is in use for live upgrade -. Please see ludelete(1M). </pre> <p> Yeah, You might think so Solaris. I move along and setup the layout. </p> <p> Then, <tt>ufsrestore</tt> the data </p> <pre> newfs /dev/rdsk/c1t0d0s0 mount /dev/rdsk/c1t0d0s0 /mnt cd /mnt ufsrestore rf /extra/holding/root.dump </pre> <p> Sensing that this might not be enough, I decide to do a comple of things. First, reinstall GRUB manually, as described in Slava's entry </p> <pre> installgrub /boot/grub/stage1 /boot/grub/stage2 /dev/rdsk/c1t0d0s0 stage1 written to partition 2 sector 0 stage2 written to partition 2, 235 sectors starting at 50 </pre> <p> And then, just for grins, update this Boot Environment to the New Boot Environment I created just a little while earlier. I figured this should ensure that anything possible that I missed with <tt>ufsdump</tt> would be restored correctly </p> <pre> lumake -n sol-nv-87 </pre> <p> Well, <tt>lumake</tt> returned success at least. Here goes nothing. Lets activate the Boot Environment and cross our fingers </p> <pre> luactivate sol-nv-87 init 6 </pre> <p> Tah-Dah! Yet another reason why I love Solaris. You flex it in weird ways and it responds logically. </p> Playing around with Apache HttpComponents mock 2008-03-21T14:45:58+00:00 2008-03-21T21:45:58+00:00 <p> The other week while I was "stuck" at home watching our little baby boy, I spent a little time ASF's <a href="">HttpComponents</a>. I have several HTTP utilities but so far I have just been using the somewhat limited <tt>java.net.\*</tt> classes. </p> <p> As a goal of the investigation, I decided to write a multi-threaded web crawler, to some level of completeness. </p> <p> Here's an overview of the heart of the crawler. </p> <p> After looking around at the HttpComponents examples, it's pretty clear that the main entry point is the <tt>org.apache.http.client.HttpClient</tt> interface. This interface represents an "HTTP Client". Analogous to say the heart of a Browser like Firefox. It mainly provides methods to allow you to execute HTTP Requests. Various subclasses exist, the main one of interest being <tt>DefaultHttpClient</tt> which has methods for setting up all the typical http goodies like Cookie stores, Authentication methods, and Connection managers. </p> <p> The simplest instantiation to take the defaults is something like </p> <pre> HttpClient httpClient = new DefaultHttpClient(); </pre> <p> But this isn't good enough for me because it creates a client that has only a single threaded Connection Manager. This will not work for my goal. A little bit more code will fix that and create an HttpClient with a multi-thread save Connection Manager. </p> <pre>); </pre> <p> What's the Connection Manager about? Well it provides more advanced connection management features, such as connection pools for things like keep alive connections. </p> <p> Ok, so now that the HttpClient is set up, I can execute HTTP requests in various ways, but one of the easiest is </p> <pre> HttpGet httpget = new HttpGet(url); HttpResponse response = httpClient.execute(httpget); HttpEntity entity = response.getEntity(); </pre> <p> With the HttpResponse and HttpEntity objects I can interrogate the status of the response, find the Content-Type, and get the content in the response. I put this all in a class I called <tt>Downloader</tt> the salient pieces are </p> <pre>) { } } } </pre> <p> My actual code is structured a little differently because I created the notion of filters to constrain the URLs that are processed - constraints like same host, same domain, exclude paths, etc. <p> <p>. </p> <p> The other curious thing you may have noticed is that I had the <tt>Downloader</tt> class implement the <tt>Runnable</tt> interface. Why is this? Well this has to do with the the multi-threading I wanted to do. It implements <tt>Runnable</tt> so I can schedule them to execute in a pool of threads. </p> <p> So how did I manage the thread pool? Why with the concurrent classes in <tt>java.util.concurrent</tt> - the <tt>ExecutorService</tt> specifically. The <tt>ExecutorService</tt> does all the hard work of managing the thread pool for you, all you really need to do is keep giving it work to do. To instantiate an <tt>ExecutorService</tt> I used </p> <pre> ExecutorService workerMgr = java.util.concurrent.Executors.newFixedThreadPool(threads); </pre> <p> And from there, giving more work to the service is as easy as </p> <pre> Downloader downloader = new Downloader(httpClient, url); workerMgr.execute(downloader); </pre> <p> After I'm done adding items, I simply wait for all the work to finish: </p> <pre> workerMgr.shutdown(); </pre> some xVM notes before I forget mock 2008-03-06T14:01:18+00:00 2008-03-06T22:01:18+00:00 <p> Its been about 6 months since I played around with xVM (then called just Xen) so I figured I'd refresh myself with Nevada Build 83. Previously I had been creating domains with <tt>xm create</tt> and pretty terse python-syntax(?) config files. Things are much easier now with <tt>virt-install</tt> </p> <p> A few notes on commands I used to create domains. </p> <p> <b>Installing Paravirtualized Fedora Core 7</b> </p> <pre> zfs create -V 8G storage/xvm-fc7-instance-1 lofiadm -a /export/xen/install/iso/F-7-x86_64-DVD.iso mount -F hsfs /dev/lofi/2 /export/xen/install/dvd/fc7-64 share /export/xen/install/dvd/fc7-64 virt-install --name fc7 --ram 1024 --paravirt --file /dev/zvol/dsk/storage/xvm-fc7-instance-1 \\ --location nfs:outpost.eng.sun.com:/export/xen/install/dvd/fc7-64 -x console=hvc0 </pre> <p> <b>Installing Paravirtualized Fedora Core 8</b> </p> <pre> zfs create -V 8G storage/xvm-fc7-instance-1 virt-install --name fc8 --ram 1024 --paravirt --file /dev/zvol/dsk/storage/xvm-fc8-instance-1 \\ --l -x console=hvc0 </pre> <p> <b>Duplicate a domain</b> </p> <pre> zfs snapshot storage/xvm-fc8-instance-1@master zfs clone storage/xvm-fc8-instance-1@master storage/xvm-fc8-instance-2 virsh dumpxml fc8-1 | grep -v uuid | sed -e 's/fc8-1/fc8-2/g' > /tmp/fc8-2 virsh define /tmp/fc8-2 </pre> <p> <b>Install HVM Solaris Install to NFS image</b> </p> <pre> virt-install -n nv-81-1 --hvm --vnc --vncport 5901 \\ -f /net/derelict/extra/holding/xen-root-nv-81-1.img -r 1024 \\ -c /export/xen/install/iso/sol-nv-bld81-x86-dvd.iso --noautoconsole </pre> Hopping on the VBox/Indiana Bandwagon mock 2008-02-16T11:07:59+00:00 2008-02-16T19:07:59+00:00 <p> Ok, so I decided to take a look at Indiana inside of VirtualBox like many others already have. The main thing people have been talking about is the lack of the <tt>pcn</tt> driver, which is needed for networking inside of VirtualBox. </p> <p> I found <a href="">Alan Burlison's Entry</a> describing how to get a copy of <tt>pcn</tt> into Indiana from an SXDE release. </p> <p> Only one problem for me, I'm at home and don't happen to have an SXDE ISO around, just a few systems with SXDE installed that I can get to. </p> <p> My first idea, use a thumb drive to copy the necessary files into - Note, I'm running VirtualBox on top of XP. So I copied the files onto the thumb drive, but ran into problems making the drive accessible inside of the Indiana Guest. </p> <p> Next, my eye caught the "Floppy Drive" options. I don't have a floppy drive in my system, you can build an image that can become a virtual floppy drive. So I hit up the ol' internet search to look for a Windows program to build a floppy image, and I found <a href="">Build Floppy Image</a>. </p> <p> A quick look at the options and it would appear to fit the bill. So I copied <tt>pcn</tt> and <tt>pcn.conf</tt> into <tt>C:\\temp</tt>, and ran </p> <pre> bfi -t=144 -f=c:\\Img\\pcn.img c:\\temp </pre> <p> Done in a flash. I configured VirtualBox to use <tt>c:\\Img\\pcn.img</tt> as Indiana's floppy. Booted it, and Voila - The floppy and files were available copy per Alan's instructions. </p> The Sys Admin Adventure Game mock 2007-11-06T16:49:55+00:00 2007-11-07T00:49:55+00:00 <p> Every once in a while I ball things up pretty good on one of my systems it takes a while to figure out how to fix things, and in the process validates that I still have at least some system administration capabilities still. </p> <p>. </p> <p> As I mentioned, I wanted to upgrade my desktop to Nevada 76, so I could take a look at <a href="">Erwann's new Solaris Build of Compiz</a>. I was currently running a hybrid build of Nevada 70 with the Xen/xVM bits in them. So I figured I would to a Live Upgrade to 76. </p> <p>. </p> <p>. </p> <p> So I attempted to <tt>pkgrm</tt> and <tt>pkgadd</tt>. </p> <p>. </p> <p> And now is where I hit the first problem. When the system boots, All I see is </p> <pre> <span style="color: red">Bad PBR Sig</span> </pre> <p> <tt>installboot</tt>. So I boot from the other disk into Failsafe mode. And try running it. </p> <pre> Error: installboot is obsolete. Use installgrub(1M) </pre> <p> Doh. Ok so I look at the manual page for <tt>installgrub</tt> and figure out the right thing to do. And attempt another boot. This time it boots GRUB only there is no GRUB menu. Simply a prompt. </p> <p> Dang! So I take a peek at another system I got running Solaris 10 to look at its <tt>/boot/grub/menu.lst</tt> file and come up with the following to run at the GRUB prompt. </p> <pre> root (hd1,1,a) kernel /platform/i86pc/multiboot module /platform/i86pc/boot_archive boot </pre> <p> I give it a try, and guess what? Well apparently <tt>multiboot</tt> is no longer supported it tells me. And suggests the following lines instead: </p> <pre> kernel$ /platform/i86pc/kernel/$ISADIR/unix module$ /platform/i86pc/$ISADIR/boot_archive </pre> <p> Sweet. Just enter these and I'm off and running, I think. Bzzt. Not yet. I do get a little farther though. The kernel does start loading but then I get some message like "Couldn't find /devices". Wha? </p> <p> Back to more web search. And I find some other random post saying the problem is because there is no <tt>bootpath</tt> in the <tt>bootenv.rc</tt>, and gives an example with a big device path name. So again I boot into Failsafe, and after I figure out the device path name, I update the <tt>bootenv.rc</tt> file and give it another go. </p> <p> Hey look at that, its starting to <tt>fsck</tt> the <tt>/usr</tt> partition. Hey wait, I have no <tt>/usr</tt> I have only a root partition. Again back to Failsafe to see what's going on. This one stumped me for a little bit, but eventually figured it out. Essentially it was becuase there was no root entry in <tt>vfstab</tt>. </p> <p> Finally fixed that up. And gave it one final reboot and Nevada finally booted into multiuser. Whew. </p> <p> 'Course that didn't fix the problem I was having with Thunderbird crashing, but it was kinda fun exercise anyways. </p> Newer, bigger, better, SLOWER! mock 2007-10-31T13:12:54+00:00 2007-10-31T20:12:54+00:00 <p> The other day I needed to put together a CRUD-type web app. In it, I knew I was going to do a bunch of XML processing, and since I only had about a week and a half to put it together, I decided to just use JSPs and the JSTL. </p> <p> I happened to have Sun Java System Web Server 6.1 already installed on one of my development boxes, so I just went ahead and started developing on it. I made heavy use of the XML processing tags in JSTL to read, iterate and transform a bunch of remote XML files. Everything went together very smoothly. </p> <p> Then came time to migrate it to a more "production" server, which happened to be running Sun Java System Web Server 7.0, and things went a little awry. The application took much longer to process all the XML files. I distilled things down to a test case where processing a 400K XML file went from about 50 milliseconds on SJSWS 6.1 to about 1/2 second with SJSWS 7.0. </p> <p> WTF! </p> <p> I knew JSTL bumped up from version 1.0 in WS 6.1 to version 1.1 in WS 7. A seemingly minor change, based on version numbers at least. </p> <p> So next I loaded the application up into Tomcat 5.5.x, just to verify it wasn't a problem with WS 7.0, which was true. The application performed just as poorly in Tomcat as WS 7. Both use JSTL 1.1. </p> <p> Well, after some digging, I found the culprit. It ends up being a problem with JSTL 1.1, and is logged as <a href="">Bug 27717</a> over at ASF. Essentially, the back end XML processing APIs were replaced between JSTL 1.0 and 1.1, from Jaxen/Saxpath to Xalan/Xerces. That's actually a pretty big difference. You can read the bug to understand it more. </p> <p> Ok, with the problem identified, I still didn't have a solution, and I wanted to use WS 7. So I thought I'd attempt to install JSTL 1.0 in WS 7. Probably not a supported configuration, but one that might just work for now. </p> <p> And this, in fact, worked just great! Pretty easy to do to. After downloading and extracting <a href="">JSTL 1.0</a>. I created a <tt>lib-ext</tt> folder at the base directory of where I installed WS 7. Then copied <tt>jaxen-full.jar, saxpath.jar, standard.jar, jstl.jar</tt> into <tt>lib-ext/</tt> </p> <p> Then I logged into the WS 7 Administration Console and navigated to <i>Configurations</i> -> <i>instance</i> -> <i>Java</i> and added the following in the <i>Classpath Prefix</i> </p> <pre> ${WS_INSTALL_ROOT}/lib-ext/jaxen-full.jar ${WS_INSTALL_ROOT}/lib-ext/saxpath.jar ${WS_INSTALL_ROOT}/lib-ext/standard.jar ${WS_INSTALL_ROOT}/lib-ext/jstl.jar </pre> <p> Then just deployed the configuration, and voila! Instant performance boost for my application. </p> Subversion over SSH Access in NetBeans on Windows mock 2007-10-09T14:47:43+00:00 2007-10-09T21:47:43+00:00 <p> The other week I described some general instructions on <a href="">How to set up Subversion over SSH on Solaris</a>. That's all well and good but to make it useful, you need to configure your clients to access your repository. In another case of "before I forget how I did it", here's some instructions on how to set up access through NetBeans running on Windows. </p> <p> If you've read any of my previous entries, you'll know that I advocate adding pass-phrases to your SSH keys. This renders them more or less useless if they fall into someone else's hands. Adding pass-phrases, however, generally presents wrinkles here and there. But, so far, they are not unsurmountable. And this case is no different. </p> <p> First off, you need something that will tunnel your Subversion access through SSH, for this, the best choice is <a href="">PuTTY</a>. You can grab an entire dist to download but all you will really need is <tt>Plink</tt>, <tt>Pagent</tt>, and <tt>PuTTYgen</tt>. So grab and install a dist that includes those. </p> <p> Now you need some SSH keys (If you remember from my last post, this whole scheme relies on SSH public-key authentication.) In my case, I already had some SSH keys in my Solaris and OS X environments, and I just wanted to migrate those keys over to PuTTY. So I grabbed the <tt>~/.ssh/id_dsa</tt> and <tt>~/.ssh/id_rsa</tt> files from my Solaris account. Then fired up <tt>PuTTYgen</tt>. </p> <p> The thing here is that PuTTY keeps its key files in a different format than SSH/OpenSSH. So you need to convert them with <tt>PuTTYgen</tt>. This is pretty easy though. Just use the <i>Conversions->Import</i> Key menu option in <tt>PuTTYgen</tt> to import your SSH/OpenSSH key, then save the converted private key with the <i>Save private key</i> option in <tt>PuTTYgen</tt>. Do this for all your private keys. Note that if you have pass-phrase on your private keys, you will be prompted for the password when you import them. </p> <p> Ok, with the keys converted, quit <tt>PuTTYgen</tt>, and fire up <tt>Pagent</tt>. The <tt>Pagent</tt> daemon is similar to the <tt>ssh-agent</tt> utility in OpenSSH. It makes your private keys available to requestors without having to re-authorize them by entering the pass-phrases. </p> <p> When you fire up <tt>Pagent</tt>, it'll put itself in the System Tray. Double click on the icon to open it up. Simply click on the <i>Add Key</i> button and select each of your keys one at a time to load. Now, just click on the <i>Close</i> button, to put it back in the System Tray. </p> <p> Oh yeah, in case you haven't already downloaded a Subversion client. Do it now. I just grabbed one of the Windows installers from the Subversion <a href="">Downloads Page</a>. </p> <p> Now, another key step, make sure both Subversion and PuTTY are in your PATH. Start up <i>Control Panel -> System</i>, then click on the <i>Advanced</i> tab, and then on <i>Environment Variables</i>. Find the Path variable in the System Variables section and add PuTTY and Subversion, in my case, "<tt>C:\\Program Files\\Subversion\\bin;C:\\Program Files\\PuTTY</tt>". </p> <p> Now, Finally, NetBeans (in my case NetBeans 6.0 Beta 1) enters the picture. Fire up NetBeans, then, the easiest thing to do is to click on <i>Versioning -> Subversion -> Checkout</i>. Enter your "svn+ssh" <i>Subversion Repository URL</i> in the appropriate field, and you will notice NetBeans will give you a hint as to what you need to enter in the <i>Tunnel Command</i> field. If you have done everything correctly, you should simply enter <tt>plink</tt>, then click on <i>Next</i>. In my case (and how I outlined in my previous entry), I need to login to the Subversion repository as the <tt>src</tt> user so I had to add an additional "<tt>-l src</tt>" option to the <tt>plink</tt> command. </p> <p> If all has gone well, when you click on <i>Next</i>, you will see NetBeans connecting to the Repository then bring you to a <i>Folders to Checkout</i> screen. Try clicking <i>Browse</i> there, and you should be able to navigate to your <tt>trunk</tt> and eventually download a working copy for your development work. </p> <p> If something has gone wrong, it can be a little tricky to figure out what the problem is. I ran the <tt>plink</tt> utility by had in a <tt>cmd</tt> window several times during initial setup to solve a few issues. </p> Setting up Subversion over SSH on Solaris mock 2007-09-26T13:29:15+00:00 2007-09-26T20:29:15+00:00 <p> Before I forget how I did it, I figured I should probably document some steps on how I set up a Subversion repository on Solaris so that it can be accessed over SSH, using svn+ssh:// URLs. </p> <p> The <a href="">Tunneling over SSH</a> section in the <b>Version Control with Subversion</b> book actually does a pretty good write up of the basics. The main tweak that I did with Solaris was to beef security slightly by creating a Solaris Role on the repository server which can only execute a limited number of subversion commands. Here's what I did. </p> <p> First off, I created a RBAC profile called "Source Code Shell" which can run only the <tt>svnserve</tt> command, and runs it as the source code management user I had previously created called <tt>scm</tt> </p> <pre> cat >> /etc/security/exec_attr Source Code Shell:suser:cmd:::/opt/svn-1.4.0/bin/svnserve:uid=scm Source Code Shell:suser:cmd:::/opt/svn/bin/svnserve:uid=scm cat >> /etc/security/prof_attr Source Code Shell:::Access Subversion Only: </pre> <p> Ok that was easy enough. There are probably some more appropriate commands to use than just adding to the files directly. I should probably look into that. </p> <p> Next, I need to create a role to assign this new profile to. The user name will be <tt>src</tt> </p> <pre> useradd -d /export/home/src -m -c "Source Code User" -s /usr/bin/pfsh -g 100 -u 242 src passwd -N src </pre> <p> And assign the RBAC Profile to the <tt>src</tt> user. </p> <pre> usermod -P "Source Code Shell" src </pre> <p> Now, the rest of this is pretty much all described in the "Tunneling over SSH" section mentioned earlier. Essentially what is done is to use the public-key authentication mechanisms in SSH to identify the incoming user and automatically start up the <tt>svnserve</tt> command in tunnel mode. </p> <p> This is accomplished by adding lines to the <tt>src</tt> user's <tt>authorized_keys</tt> file for each user who will be accessing the repository. The one thing you need for each user is their public-key file(s), typically <tt>id_dsa.pub</tt> or <tt>id_rsa.pub</tt> The format of the lines in <tt>authorized_keys</tt> is </p> <pre>NASCAR</a> "fantasy" pool with some friends and family. One of the features I have is a live update mechanism to see where everyone stands during the race. </p> <p> Only problem, I couldn't find a simple, easy data feed to get the current race information. Ok, so, the next best thing, look at one of the sports sites like the <a href="">Yahoo NASCAR update</a> and try to extract information out of the HTML. </p> <p> Well, the data looks relatively well formatted, but as we all know browsers are pretty lenient about what they accept as HTML, what with missing close tags and so on, and I wanted to use an XML parser, and even XSLT to extract the data. So I needed a way to fix the HTML before passing it to a parser. </p> <p> Enter <a href="">Tagsoup</a>. A SAX-compliant HTML parser that spits out well-formatted XML. Ah that sounds like just the ticket. And even better, the maintainers include a modified version of Saxon - <a href="">TSaxon</a> - to process XSLT. </p> <p>So with TSaxon, in hand, it made easy work of converting something like an <a href="">ESPN Qualifying Grid</a> into SQL that I can load into Derby with an XSL like <pre> <xsl:stylesheet xmlns: <xsl:param <xsl:param <xsl:param <xsl:param <xsl:template <xsl:apply-templates </xsl:template> <xsl:template <xsl:text>delete from </xsl:text> <xsl:value-of <xsl:text> where season = </xsl:text> <xsl:value-of <xsl:text> and race = </xsl:text> <xsl:value-of <xsl:text>; </xsl:text> <xsl:apply-templates <xsl:text>commit; </xsl:text> </xsl:template> <xsl:template <xsl:variable <xsl:variable <xsl:variable <xsl:text>insert into </xsl:text> <xsl:value-of <xsl:text> values (<>); </xsl:text> </xsl:template> </xsl:stylesheet> </pre> <p> By passing it to TSaxon like </p> <pre> java -jar lib/saxon.jar -H "$1" xsl/grid-to-sql.xsl race=$2 season=$3 table=grid type=$4 </pre> <p> Ok, so that's pretty cool by itself. But now with the NFL season just beginning, and another private "fantasy" pool among relatives, I found that I wanted to do a similar "live tracker" to see everyone's points in the pool. This time I wanted to do it a little bit differently. </p> <p> I wanted to programmatically call Tagsoup to parse a page and pass it through an XSLT. So why not try to use all the JAXP facilities in Java. And it turns out to be pretty easy. This easy <pre> // Create an instance of Tagsoup SAXBuilder builder = new SAXBuilder("org.ccil.cowan.tagsoup.Parser"); // Parse my (HTML) URL into a well-formed document Document doc = builder.build(new URL(poolurl)); JDOMResult result = new JDOMResult(); JDOMSource source = new JDOMSource(doc); // Get a JAXP Factory TransformerFactory factory = TransformerFactory.newInstance(); // Get a Transformer with the XSL loaded. StreamSource sheet = new StreamSource(sheetpath); Transformer transformer = factory.newTransformer(sheet); // Transform the page transformer.transform(source, result); // Spit out the result. XMLOutputter outputter = new XMLOutputter(Format.getPrettyFormat()); outputter.output(result.getDocument(), out); </pre> <p> The middle section is really the part where JAXP comes into play. But as you can see its quite simple. </p> <p> Oh, and say you want to just programmatically select nodes with XPath. That's pretty easy too. Here's an example that gets the Title of the page </p> <pre> JDOMXPath titlePath = new JDOMXPath("/h:html/h:head/h:title"); titlePath.addNamespace("h", ""); String title = ((Element) titlePath.selectSingleNode(doc)).getText(); out.println("Title is " + title); </pre> <p> Voila. </p> Updated JVM Options List mock 2007-08-28T15:13:33+00:00 2007-08-28T23:00:54+00:00 <p> Its been quite a while, but I finally spent some time to update my <a href="">List of JVM Options</a> to include everything in Java 6. This is just the first pass as Java 6, so most of the new Java 6 options have no description for them. I still need to go out and search for references for any/all of them. </p> <p> Hope people find this useful. </p> sun.com performance mock 2007-04-16T15:05:33+00:00 2007-04-16T22:05:33+00:00 <p> A couple of months ago, we went through a tuning exercise for our primary web site <a href=""></a>. A few execs noticed some strange pauses during page rendering (of primary the home page.) So a few of us started to look closer at the problem. </p> <p> Before I get into it a bit, I just took a look at the daily reports again for today, and I'm happy to report that performance is still looking pretty good. <p> <p> Although I'm not currently on the engineering team for the site, the I was one of the primary "architects" for the framework that serves the site, and I still have soft spot for it. I take offense when someone challenges the performance of the site. So, I decided to get involved, mainly to help gather some statistics for how the site is performing. </p> <p> So what did performance look like when we looked into it? Here are a couple of charts showing what it looked like. The first chart shows the average response time to retrieve all the home page components. The second shows each component in detail. </p> <center> <table border="0" width="60%"> <tr> <td align="center"> <a href="/mock/resource/tuning-1/avg-dur-2007-02-05-sun.png"><img src="/mock/resource/tuning-1/avg-dur-2007-02-05-sun-sm.png"></a> <br> Averages </td> <td align="center"> <a href="/mock/resource/tuning-1/dur-2007-02-05-sun.png"><img src="/mock/resource/tuning-1/dur-2007-02-05-sun-sm.png"></a> <br> Details </td> </tr> </table> </center> <p> The charts don't look all that great. Very erratic, and pretty poor average response time. Now, the home page has about 70 components totaling about 396K. Making a very rudimentary guess at a theoretical minimum download time of all the components over say a 1.5Mbit DSL line, we're looking at about 2 seconds [ ((396 \* 8 / 1024) / 1.5) = 2 seconds ]. Now lets compare that against what we see on the average duration chart. Assuming a sequential download of components, even a 100 millisecond response time would take 7 seconds to download all the comonents [ 100 \* 70 / 1000 = 7 ]. Which wasn't far from the truth. </p> <p> One thing to note, through our analysis, there is actually some parallelism going on to retrieving the page components. There is a great Firefox Extension that will show this visually called <a href="">Firebug</a>. So that helps reduce some overall download time. </p> <p> Eventually we found the culprit. Having narrowed the problem down to some thing with Sun Java System Web Server, <a href="">Chris Elving</a> with Web Server Engineering helped us resolve the problem. As it turned out, it ended up being <a href="">Reverse Proxy Plugin Bug # 6435723</a>, which, fortunately had a fix available for it. </p> <p> So, how are we looking today? Well, here's the charts as of Friday. Not much fluctuation in response times throughout the day. </p> <center> <table border="0" width="60%"> <tr> <td align="center"> <a href="/mock/resource/tuning-1/avg-dur-2007-04-13-sun.png"><img src="/mock/resource/tuning-1/avg-dur-2007-04-13-sun-sm.png"></a> <br> Averages </td> <td align="center"> <a href="/mock/resource/tuning-1/dur-2007-04-13-sun.png"><img src="/mock/resource/tuning-1/dur-2007-04-13-sun-sm.png"></a> <br> Details </td> </tr> </table> </center> <p> I'm sure you notice the occasional blips of response times in the over 1.5 second range. Well this is due to occasional cache flushes and page rendering due to content changes. Our rendering framework makes extensive use of XML and J2EE technologies to dynamically render the web site. I'd say in light of that, the platform is doing an incredible job at serving content. </p> <p> For comparison... Here's similar charts for one of our competitors from Friday as well. Their performance hasn't changed much over the last two months. I'm not gonna tell them. </p> <center> <table border="0" width="60%"> <tr> <td align="center"> <a href="/mock/resource/tuning-1/avg-dur-2007-04-13-comp.png"><img src="/mock/resource/tuning-1/avg-dur-2007-04-13-comp-sm.png"></a> <br> Averages </td> <td align="center"> <a href="/mock/resource/tuning-1/dur-2007-04-13-comp.png"><img src="/mock/resource/tuning-1/dur-2007-04-13-comp-sm.png"></a> <br> Details </td> </tr> </table> </center> And now Chicken of the VNC tunneled through SSH on OS X mock 2007-03-22T15:19:17+00:00 2007-03-22T23:19:17+00:00 <p> After I posted yesterday about the <a href="">VNC Over SSH Startup Script</a> I use with TightVNC on Solaris, I started thinking about how I could do the same thing with my MacBook Pro and <a href="">Chicken of the VNC</a>. </p> <p> And I have just completed the first version. Its amazing what you can fumble through with an internet search engine. So here we go... </p> <p> <b>Password Please</b> -- One thing you don't get with CotVNC is the <tt>vncpasswd</tt> command. So we need a copy. I just did it the quick and dirty way by grabbing the TightVNC source and compiling only the <tt>vncpasswd</tt> command: </p> <pre> </pre> <p> Now squirrel off the <tt>vncpasswd/vncpasswd</tt> command in your favorite <tt>bin/</tt> directory. </p> <p> <b>Passphrase Please</b> -- If you look back at my previous post, I make use of <tt>ssh-agent</tt> and <tt>ssh-add</tt> to allow tunneling without entering a password. Only problem on OS X is that there is no <tt>gnome-ssh-askpass</tt> command available. So I had to hack one up myself. Its an odd combination of Bourne Shell and AppleScript, but it appears to work. Save a copy as <tt>macos-askpass</tt>. </p> <pre> #! /bin/sh # # An SSH_ASKPASS command for MacOS X # # Author: Joseph Mocker, Sun Microsystems # # To use this script: # setenv SSH_ASKPASS "macos-askpass" # setenv DISPLAY ":0" # TITLE=${MACOS_ASKPASS_TITLE:-"SSH"}NFL Pool</a> with my extended family last season, my wife and I started thinking about how we could run a simple NASCAR Nextel Cup Series pool. We didn't want to go whole hog fantasy league or nothing, just something simple. </p> <p> What we decided on was a simple pool where you picked the top 10 drivers to complete the race, and you got points for what position they came in relative to your picks. </p> <p> As we talked it out, I started thinking about how I could manage something like this. A database seemed natural for managing the information. So I worked out a little schema, and decided to play with the <a href="">Apache Derby</a> database. </p> <p>. </p> <p> But it didn't stop there. I started thinking about the <a href="">Fanpool</a>. </p> <p>. </p> . </p> <p> About this time, I had one of my brother's in law playing with us. And my father too. So Of course now with four people playing, I needed to create a authentication mechanism, and a way for them to enter their picks themselves. </p> <p> So you can see where this is going. It just ballooned from there. Now I even have it set up so it can display results realtime by pulling down some stats from <a href="">ESPN</a> during the race. </p> <p> Its all been kinda fun to do. And now we even have 8 or so people playing in our pool. And since a few people started just last week, I even threw in the ability to spot people points. </p> <p> Anyways, if you want to check out a demo, go to <a href="">The NASCAR Pool</a> and check out the demo pool I have set up. It makes the Nextel Cup pretty fun to watch when you throw some personal competition into it. </p> Compiling UW imapd with SSL on Solaris 10 mock 2007-03-14T11:34:59+00:00 2007-03-14T19:34:59+00:00 <p> I've been running IMAP over SSL for a while on Solaris, but until recently I've used <a href="">STunnel</a> to provide the SSL support in front of a plain IMAP daemon. I've known that you could compile SSL into imap for a while but never really looked into it until <a href="/rama">Rama</a> figured out the magic certificate generation piece. </p> <p> But what Rama did was to just install the <a href="">Sunfreeware</a> version of imapd. I have a love/hate relationship with those types of distributions, so I decided to look at compiling it myself. Heck Solaris includes OpenSSL so it should be easy. </p> <p> Well, actually, I couldn't get it to build with the version of OpenSSL that ships with Solaris. Looking at <tt>syslog</tt> I'd see messages like: </p> <pre> Mar 14 10:23:24 watt imapd[5834]: [ID 853321 mail.error] SSL error status: error:140D308A:SSL routines:TLS1_SETUP_KEY_BLOCK:cipher or hash unavailable </pre> <p> And looking at the imapd binary I saw a missing <tt>libcrypto_extra</tt>. Searching the net I saw a bunch of people talking about it. It appears that this is no longer needed with Solaris 10, but others say that you need to install {{SUNWcry}} package. Well, I must be a loser because I could not find enough info to make it work. </p> <p> So I decided to just compile up a fresh copy of OpenSSL to use to compile imapd. So here's what I did. </p> <p> <b>Compiling OpenSSL</b> -- its pretty trivial to do, in this day and age, however my first attempt compiled it 64bit, and imapd had issues with that. There are a few extra configuration parameters to force it to 32 bit. Here's the <tt>Configure</tt> line. </p> <pre> Configure --prefix=/opt/openssl-0.9.8e 386 shared solaris-x86-gcc </pre> <p> After that.. Compile and Install... </p> <pre> gmake gmake install </pre> <p> <b>Compiling Imapd</b> -- The instructions in <tt>docs/SSLBUILD</tt> go over the basics. But there were a few additional changes I needed to make. The main change was to make sure imap was built with my OpenSSL instead of the Solaris version. All these changes were to <tt>src/osdep/unix/Makefile</tt>: </p> <p> Fist I set the <tt>SSLDIR</tt> and <tt>SSLCERTS</tt> variables to where I wanted them: </p> <pre> SSLDIR=/opt/openssl SSLCERTS=/etc/sfw/openssl/certs </pre> <p> Next, I forced it to use the static version of libcrypto.a by changing <tt>SSLCRYPTO</tt>: </p> <pre> SSLCRYPTO=$(SSLLIB)/libcrypto.a </pre> <p> Finally, I need to force it to use my static version of libssl.a. </p> <pre> SSLLDFLAGS= -L$(SSLLIB) $(SSLLIB)/libssl.a $(SSLCRYPTO) $(SSLRSA) </pre> <p> After that. Simply compile it up, and install it where ever you want: </p> <pre> gmake gso mkdir /opt/bin cp imapd/imapd /opt/bin </pre> <p> <b>Configuring the imapd certificate</b> -- Thanks to <a href="/rama">Rama</a> on the magic OpenSSL command. All that you really do is create a PEM certificate called <tt>imapd.pem</tt> in the OpenSSL <tt>certs</tt> folder: </p> <pre> cd /opt/sfw/openssl/certs openssl req -new -x509 -nodes -out imapd.pem -keyout imapd.pem -days 3650 </pre> <p> <b>Starting imapd from inetd</b> -- Ok well now with Solaris 10 this is done though SMF, but inetd has a conversion utility to do this. I put the following line in <tt>/etc/inetd.conf</tt> </p> <pre> imaps stream tcp nowait root /opt/bin/imapd imapd </pre> <p> Then added a line to <tt>/etc/services</tt> </p> <pre> imaps 1143/tcp imap2 # Internet Mail Access Protocol v2 </pre> <p> Then just run <tt>inetconv</tt> per instructions in <tt>inetd.conf</tt> and bob's your uncle. </p> One of my favorite Firefox extensions mock 2007-03-13T22:27:56+00:00 2007-03-14T06:27:56+00:00 <p> I use several computers on a daily basis. I've done this for quite a while. At work I have my workstation, at home I have a PC and lots of times I just use my laptop. The problem is I like to try to keep my desktop environment in sync across all those machines as much as possible. </p> <p> With Firefox, for example, I want all my bookmarks up to date everywhere. And fortunately, for Firefox, I have found an extension quite some time ago which helps do this. </p> <p> Its currently called <a href="">Bookmark Sync and Sort</a>, but previously it was called <i>Bookmark Synchronizer [2,3]</i>. </p> <p> What this extension allows you to do is to store your bookmarks on server somewhere, and load them in multiple Firefox instances. It does some cool things like allowing you to merge bookmarks on your server with any changes you've made locally. And it provides multiple protocols for saving and loading the bookmarks, including FTP and WebDAV. </p> <p> What I did was set up Sun Java System Web Server 7.0 on a server I can access from anywhere I am, and configured a WebDAV collection for <i>Bookmark Sync and Sort</i>. What about security? Well a couple of things. My SJSWS instance is configured for SSL with a self-signed certificate. And I have restricted the WebDAV collection to authenticated users. <i>Bookmark Sync and Sort</i> handles all this great. </p> <p> So how does one configure WebDAV with SJSWS 7.0? Its actually not that difficult. A good reference is <a href="">Meena Vyas' SJSWS WebDAV entry</a>. <p> <p> Specifically, here's what I did. First enable WebDAV with wadm </p> <pre> wadm> enable-webdav --config=<server-instance> </pre> <p> Then create a WebDAV collection </p> <pre> wadm> create-dav-collection --config=<server-instance> --vs=<virtual-server> --uri=/davx/marks --source-uri=/marks </pre> <p> This will create a WebDAV collection that you access via DAV at <tt>/marks</tt>, which is a little opposite of usual with WebDAV but we want to make sure the web server doesn't mess with the bookmarks file when it serves it. </p> <p> Now, that is enough to set up an open WebDAV collection. But it would be good to control access to the collection. So, from the Admin Console, Set up access control. Go to the <i>Access Control</i> tab of <i>Configurations -> server-instance</i>. </p> <p> Then select the <i>Users</i> tab and create or edit the users and passwords. </p> <p> Next, go to the <i>Access Control Lists (ACL)</i> tab. Edit the <b>default</b> ACL, and add an entry with the settings </p> <ul> <li> Access: Allow <li> Users & Groups: All in the authentication database <li> Rights: All Access Rights </ul> <p> Next, Edit the <b>dav-src</b> ACL, add a similar entry with the settings <p> <ul> <li> Access: Allow <li> Users & Groups: All in the authentication database <li> Rights: All Access Rights </ul> <p> Now, Deploy the configuration and you are done. </p> Force SSL Web Server Hack mock 2006-11-23T10:31:01+00:00 2006-11-23T18:31:01+00:00 <p> I've finally gotten around to upgrading my personal website from Sun Java System Web Server 6.1 to 7.0 - as well as a refresh of <a href="">JSPWiki</a> <p> <p> As I have been doing the migration, one of the things I ran across was a method I have for forcing a URLs to the SSL/HTTPS port of the web server. For example, I have a <a href="">WebMail</a> application installed that I occasionally use. The web server is configured to serve both Plain HTTP and SSL HTTP requests. But in this case I don't want to be able to access the WebMail application over plain HTTP. How do I force SJSWS to redirect any plain HTTP requests to the SSL? Its really pretty easy. </p> <p> This works for both 6.1 and 7.0. All you really do is edit the <tt>obj.conf</tt> (or with 7.0, it might be <tt>node-obj.conf</tt>) and add a <tt>Client</tt> tag to the beginning of the "default" <tt>Object</tt> of the form: </p> <pre> <Client match="all" security="false"> NameTrans NameTransDeploying Wikis to Sun Java System Web Server 7.0, Part 1: JSPWiki</a> reminded me that I wanted to write a few things about my experiences with deploying JSPWiki on SJSWS 7.0 </p> <p> The article does a good job at explaining the basics of installation and deployment. SJSWS 7 really makes it pretty easy to deploy web applications. </p> <p> The tougher part is how to configure the JAAS components in JSPWiki with SJSWS. Actually once you get your bearing with SJSWS, its not really that difficult at all. There are two components to JAAS that you need to deal with. These are the <i>JAAS login configuration</i> and the <i>Java 2 security policy</i>. And here is how to do it. </p> <h3>JAAS login configuration</h3> <p> SJSWS comes with a login configuration already installed. The trick is to find the configuration file and add the JSPWiki configuration to it. The file is called <tt>login.conf</tt> and it is located in the <tt>config/</tt> folder of the web server instance. The contents look something like </p> <pre> fileRealm { com.iplanet.ias.security.auth.login.FileLoginModule required; }; ldapRealm { com.iplanet.ias.security.auth.login.LDAPLoginModule required; }; solarisRealm { com.iplanet.ias.security.auth.login.SolarisLoginModule required; }; nativeRealm { com.iplanet.ias.security.auth.login.NativeLoginModule required; }; </pre> <p> What needs to be done is simply add the JSPWiki configuration to the end of the file. The JSPWiki configuration typically looks like </p> <pre>; }; </pre> <p> Just add those lines to the end of <tt>login.conf</tt> and you are done. </p> <h3>Java 2 security policy</h3> <p> JSPWiki uses a standard Java 2 security policy to control access to one or more JSPWiki instances within a servlet container. The policy file comes with JSPWiki is <tt>WEB-INF/jspwiki.policy</tt>. </p> <p> Now, in my observation, JSPWiki will find this file without doing anything, so this part is somewhat optional. However, without doing any explicit configration you will generally see the error messages: </p> <pre> WARN com.ecyrd.jspwiki.auth.PolicyLoader - You have set your 'java.security.policy' to point at '/opt/sun/SUNwbsvr7/ https-node/config/file:/opt/sun/SUNwbsvr7/https-node/web-app/node/wiki/WEB-INF/jspwiki.policy', but that file d oes not seem to exist. I'll continue anyway, since this may be something specific to your servlet container. Just c onsider yourself warned. WARN com.ecyrd.jspwiki.auth.PolicyLoader - I could not locate the JSPWiki keystore ('jspwiki.jks') in the same direc tory as your jspwiki.policy file. On many servlet containers, such as Tomcat, this needs to be done. If you keep hav ing access right permissions, please try copying your WEB-INF/jspwiki.jks to /opt/sun/SUNwbsvr7/https-node/config/fil e:/opt/sun/SUNwbsvr7/https-node/web-app/node/wiki/WEB-INF </pre> <p> The bit about "Consider yourself warned" is a little dubious. And the fix is pretty easy. So why not just do it. </p> <p> What needs to be done is to explicitly set the property <tt>java.security.policy</tt> for the container to the location of the <tt>jspwiki.policy</tt> file. This can be done easily in the administration console. Simply log in to the administration console and drill down into <i>Configurations -> node -> Java</i>, then click on <i>JVM Settings</i>. </p> <p> The first section you will see on this page is a <i>JVM Options</i> table, which will include, among other things, a reference to the <tt>login.conf</tt> file that we tweaked earlier. Simply click on <i>New</i> to add another value, and enter a value like </p> <pre> -Djava.security.policy=/opt/sun/SUNwbsvr7/https-node/web-app/node/wiki/WEB-INF/jspwiki.policy </pre> <p> Then click <i>Save</i> and Deploy your configuartion as described in Sun Developer Network article and you are golden. </p> Upgrading Bit Torrent mock 2006-07-13T16:02:36+00:00 2006-07-13T23:02:36+00:00 <p> The week before the break, Alon R at Azureus Inc. came in to help us redo our Bit Torrent servers. It all went pretty well. Working together it took us only a couple of days to implement a more simplified seeder and tracker network at Sun. </p> <p>The main goal that we were trying to accomplish was to simplify authoring/management of torrents. I have described the <a href="">original implementation</a> a while ago. The authoring piece was a mess, because it required remote displaying Azureus from a remote data center. All in all the process was perhaps 15 steps. Not something that could be handed off to a support team very easily. </p> <p>The other major requirements of the environment included</p> <ul> <li>High Availability - both the seeders and more importantly the tracker needed to run on multiple machines and fail over automatically.</li> <li>Geographical Location Screening - (aka rDNS) we needed to make sure that people from embargoed countries were denied access.</li> <li>Simple remote administration - the most preferable would be the web</li> </ul> <p> All in all it went together pretty quickly. I threw a couple of curves at Alon, most notably, being able to fail over the tracker to a secondary box, but he was able to get patches from the developers overnight to allow for it. </p> <p> We did encounter a strange problem when we attempted to fire up Azureus on the first server, we received a ton of Exceptions with stacktraces of the form </p> <pre> java.io.IOException: Invalid argument at sun.nio.ch.DevPollArrayWrapper.poll0(Native Method) at sun.nio.ch.DevPollArrayWrapper.poll(DevPollArrayWrapper.java:158) at sun.nio.ch.DevPollSelectorImpl.doSelect(DevPollSelectorImpl.java:68) at sun.nio.ch.SelectorImpl.lockAndDoSelect(SelectorImpl.java:69) at sun.nio.ch.SelectorImpl.select(SelectorImpl.java:80) at sun.nio.ch.SelectorImpl.select(SelectorImpl.java:84) </pre> <p> We we're both perplexed. I know I had run Azureus on Solaris 10 before and not seen the problem. And it was a new one to Alon as well. I decided to look on <a href="">bugs.sun.com</a> for lack of anything else to try. Low and behold, I found the problem. The bug <a href=""><tt>6322825</tt></a> <tt>Selector fails with invalid argument on Solaris 10</tt>. And the bug had a work around. Just set the maximum number of file descriptors on the process over 8K. Once I did that, it fired up like a charm. </p> <p> After that, the only other major task was to hook into our rDNS provider. From my description, Alon thought we could write a plugin, and again, talked to the developers who provide us a shell of a plugin. All I needed to do was add about 10-20 lines of code which actually did the screening query with the provider. </p> <p> In the end, the network consists of three servers, a Master Tracker and two Seeders. </p> <p> The Master Tracker is configured to watch a particular folder for new and removed files. When it finds changes, it automatically creates torrents for the files and starts seeding them. It also copies the <tt>.torrent</tt> files to a separate folder within the web servers docroot for customers to access. </p> <p> The two Seeders, watch an RSS feed provided by the Master Tracker. The RSS feed contains a list of all the torrents the Master is seeding. When the seeders detect a change to the RSS feed, they act appropriately. If its a new torrent, they download it from the Master and start seeding it. If if a torrent is removed, they remove it from themselves as well. </p> <p> Now, authoring a torrent is a easy a rsync'ing, scp'ing or sftp'ing a file over to the Master Seeder and voila. </p> Slightly painful migration to ZFS mock 2006-07-13T15:17:36+00:00 2006-07-13T22:17:36+00:00 <p> Fresh back from the break I decided to give and upgrade to Solaris 10_U2 a go, and migrate my data to ZFS (now included w/ U2). </p> <p> The ZFS migration was slightly painful so I figured I'd post my experience, in case anyone else might want to attempt it. </p> <p> Actually migrating the data itself to a ZFS partition was simple, however I wanted to mirror the data as I have been doing with UFS/SVM which caused the problem. Apparently logged as a known bug: </p> <pre> 6355416 zpool scrubbing consumes all memory, system hung </pre> <p> Even though the system was unresponsive, I decided to let it do its thing overnight and it did eventually finish resyncing/mirroring, and now everything is fine. </p> <P>So for what its worth, here it is. The system involved was a SunBlade 2000 with 3G of RAM. <p>The goal was to combine three separate UFS partitions (<tt>/app, /work, /extra</tt>) into a single ZFS pool which would then host all those partitions again. The partitions would then use portions of the pool as need be, precluding the need to set arbitrary sizes as has been needed with UFS. </p> <p>Also, the partitions were mirrored, and I wanted to continue mirroring with ZFS. </p> <p>The three partitions are </p> <pre> /dev/md/dsk/d4 90030867 30328334 58802225 35% /extra /dev/md/dsk/d5 24795240 23208057 1339231 95% /app /dev/md/dsk/d6 20646121 16648727 3790933 82% /work </pre> <p>(Notice that I was bumping up against some partition limits for a couple of the filesystems. The primary reason I wanted to convert to ZFS was to just have a big pool of space which could be used by whoever needed it.) <p>With the following SVM meta partitions assigned </p> <pre> d4 had submirrors d40 (c1t1d0s4), d41 (c1t2d0s4) d5 had submirrors d50 (c1t1d0s5), d51 (c1t2d0s5) d6 had submirrors d60 (c1t1d0s6), d61 (c1t2d0s6) </pre> <p>The method I went through was </p> <ul> <li>break the mirrors on one disk </li> <li>combine the partitions </li> <li>create a ZFS pool from the combined partition </li> <li>migrate the UFS/SVM data to the ZFS pool </li> <li>delete the remaining UFS/SVM partitions and combine them </li> <li>attach the combined partition as the second side of the mirror </li> </ul> <p>Here's the procedure. </p> <p><b>Break the mirrors on one disk</b> </p> <pre> metadetach d4 d41 metaclear d41 metadetach d5 d51 metaclear d51 metadetach d6 d61 metaclear d61 </pre> <p><b>Remove the meta db from the disk</b> </p> <pre> metadb -d /dev/dsk/c1t2d0s7 </pre> <p><b>Combine the disk partitions</b> </p> <p>I used the <tt>partition</tt> commands in the <tt>format</tt> utility to change the c1t2d0 disk. </p> <p><b>Create a ZFS pool called <tt>storage</tt> and filesystems within it</b> </p> <pre> zpool create storage c1t2d0s4 zfs create storage/extra zfs create storage/work zfs create storage/app </pre> <p><b>Migrate the UFS/SVM data to the ZFS filesystems</b> </p> <pre> cd /work find . -depth -print | cpio -pdmv /storage/work cd /app find . -depth -print | cpio -pdmv /storage/app cd /extra find . -depth -print | cpio -pdmv /storage/extra </pre> <p><b>Unmount the UFS/SVM partitions</b> </p> <pre> unshareall umount /app umount /work umount /extra </pre> <p><b>Remount the ZFS partitions where I expect them</b> </p> <pre> zfs set mountpoint=/work storage/work zfs set mountpoint=/extra storage/extra zfs set mountpoint=/app storage/app shareall </pre> <p><b>Remove the remaining SVM meta partitions</b> </p> <pre> metaclear d4 d40 metaclear d5 d50 metaclear d6 d60 </pre> <p><b>Remove the meta db from the disk</b> </p> <pre> metadb -d /dev/dsk/c1t1d0s7 </pre> <p><b>Combine the disk partitions</b> </p> <p>I used the <tt>partition</tt> commands in the <tt>format</tt> utility to change the c1t1d0 disk. </p> <p><b>Finally, add the combined partition as a ZFS mirror</b> </p> <p>This was the painful step. Apparently ZFS decided to take all available CPU and memory in the system. The system became unresponsive. Apparently ZFS is a little to aggressive with its resyncing (they call it scrubbing/resilvering). </p> <p>I found bug <a class="external" href="">6355416</a><img class="outlink" src="images/out.png" alt="" /> which appears to describe the issue. </p> <p>I would recommend before issuing this last command, that you boot the system into single user mode, and kill and processes that you don't need. </p> <pre> zpool attach storage c1t2d0s4 c1t1d0s4 </pre> <p>The system will probably become unresponsive within a few minutes. If so, just walk away and let it do its thing. You can use the <tt>zpool status</tt> command to check the progress of the resilver. </p> | http://blogs.oracle.com/mock/feed/entries/atom | CC-MAIN-2015-22 | refinedweb | 17,312 | 62.38 |
Created on 2018-04-21 19:10 by mcepl, last changed 2020-07-03 22:01 by mcepl.
I am in the process of writing script working with IMAP first time using Python 3 for it (unfortunately, most of servers where I run other code is so ancient that even Python 2.7 is a stretch), and it is really nice experience so far. Many problems which are dealt with on the StackExchange with arcane workarounds are now resolved in email or imaplib libraries. Thank you, everybody!
However, it seems to me that few higher level commands could greatly improve useability of imaplib library. For example, moving messages (not fault of imaplib) is still horrible mess. In the end I had to write this monstrosity just to make moving messages working:
Capas = collections.namedtuple('Capas', ['MOVE', 'UIDPLUS'])
def __login(self, host='localhost', username=None, password=None, ssl=None):
self.box = imaplib.IMAP4_SSL(host=host)
ok, data = self.box.login(username, password)
if ok != 'OK':
raise IOError('Cannot login with credentials %s' %
str((host, username, password,)))
ok, data = box.capability()
capas = data[0].decode()
box.features_present = Capas._make(['MOVE' in capas, 'UIDPLUS' in capas])
def move_messages(self, target, messages):
if self.box.features_present.MOVE:
ok, data = self.box.uid('MOVE', '%s %s' % (messages, target))
if ok != 'OK':
raise IOError('Cannot move messages to folder %s' % target)
elif self.box.features_present.UIDPLUS:
ok, data = self.box.uid('COPY', '%s %s' % (messages, target))
if ok != 'OK':
raise IOError('Cannot copy messages to folder %s' % target)
ok, data = self.box.uid('STORE',
r'+FLAGS.SILENT (\DELETED) %s' % messages)
if ok != 'OK':
raise IOError('Cannot delete messages.')
ok, data = self.box.uid('EXPUNGE', messages)
if ok != 'OK':
raise IOError('Cannot expunge messages.')
else:
ok, data = self.box.uid('COPY', '%s %s' % (messages, target))
if ok != 'OK':
raise IOError('Cannot copy messages to folder %s' % target)
ok, data = self.box.uid('STORE',
r'+FLAGS.SILENT (\DELETED) %s' % messages)
if ok != 'OK':
raise IOError('Cannot delete messages.')
It would be nice if some capabilities detection (see issue 18921) was embedded into login method, and if some version of move_messages was included in imaplib itself.
See also bpo-33336: [imaplib] MOVE is a legal command.
If I'm not mistaken, this is applied to the openSUSE TW version of Python.
For some reason, this seems to not play well with <instance>.uid('move',...)
on a cyrus imap server (v2.4.19). Is that to be expected?
```
2020-07-03 18:04:05 INFO: [imap_reorg] move b'10399' from 2017-01-01 06:30:35+02:00 to INBOX.2017
Traceback (most recent call last):
File "./imap_reorg.py", line 431, in <module>
sys.exit(main())
File "./imap_reorg.py", line 425, in main
return process()
File "./imap_reorg.py", line 133, in trace_and_call
result = func(*args, **kwargs)
File "./imap_reorg.py", line 358, in process
ret |= reorg.run_expr(expr)
File "./imap_reorg.py", line 345, in run_expr
return method(*args)
File "./imap_reorg.py", line 328, in yearly
ret = self.imap.uid('move', uid, dest)
File "/usr/lib64/python3.8/imaplib.py", line 881, in uid
typ, dat = self._simple_command(name, command, *args)
File "/usr/lib64/python3.8/imaplib.py", line 1205, in _simple_command
return self._command_complete(name, self._command(name, *args))
File "/usr/lib64/python3.8/imaplib.py", line 1030, in _command_complete
raise self.error('%s command error: %s %s' % (name, typ, data))
imaplib.error: UID command error: BAD [b'Unrecognized UID subcommand']
```
1. no this has not been included anywhere, just in the unfinished PR on GitHub
2. only thing which I was fighting to get into Python (and I did) was but that’s another issue (without this whole discussion here would not be even possible), it is now part of the upstream Python
3. are you certain that the IMAP server in question supports MOVE command? (have you noticed all that business with CAPABILITIES and particularly the UID one?)
Sorry, UIDPLUS capability. | https://bugs.python.org/issue33327 | CC-MAIN-2020-40 | refinedweb | 652 | 62.04 |
.
Note: attr_readonly was either buggy or not exposed prior to 1.2.3. If you are using a version of rails prior to 1.2.3 you can do this instead:
def attributes_with_quotes(include_primary_key = true) attributes.inject({}) do |quoted, (name, value)| if column = column_for_attribute(name) # original: # quoted[name] = quote_value(value, column) unless !include_primary_key && column.primary quoted[name] = quote_value(value, column) unless !include_primary_key && (column.primary || ["your_attributes", "listed_here"].include?(column.name)) end quoted end end.
We used something like #2 on a current project. But because the legacy database was in active use on a live site and was only periodically ported over to the new app and the schema of the app was evolving we decided to create an extra stream of migrations for the legacy import using this:
May 10, 2008 at 4:32 am
I had to migrate a M$SQL DB to a MySQL DB about a year ago. Not finding a suitable driver for M$SQL at the time for AR, I did the following:
1. Wrote an ETL process in Java that dumped date from M$SQL into a MySQL tmp schema, preserving the current schema structure.
2. Exported the current data into yml, doing all the transformations I wanted at this point.
3. The DB was small enough that I actually didn’t dump out to yml files but held the info in memory. Using key_list and value_list from fixtures.rb (I had to slightly adapt them) I wrote a small method to execute insert statements (assume objects is your yml hash). I ran into some trouble trying to import directly with yml but honestly cannot remember what it was.
objects.each {|key, value|
ActiveRecord::Base.connection.execute “INSERT INTO #{table_name} (#{key_list(value)}) VALUES (#{value_list(value)})”
}
Hope it helps,
Scott
May 12, 2008 at 3:58 am | http://pivotallabs.com/standup-5-9-2008/ | CC-MAIN-2013-20 | refinedweb | 302 | 56.05 |
REST API on Android Made Simple or: How I Learned to Stop Worrying and Love the RxJava dealing with it. Combine that with tedious work of manually dealing with any kind of RESTful API, and you soon find yourself in a hell made of disjointed code fragments, repeated loops and confusing callbacks.
This article will show you, step-by-step, how to make a simple API call on Android the easy way — using 3rd party libraries. In the first part we’ll deal with the API call itself using Retrofit and Gson. In the second part we’ll see how we can make our lives even simpler and deal with asynchronicity in a concise and elegant way using RxJava.
Disclaimer — This article does not cover the subject of RESTful APIs in great detail. It is not an in-depth Retrofit or RxJava tutorial. Rather, it’s a short and simple guide on dealing with an API request using said two libraries in conjunction.
Part I: Making API calls with Retrofit
For this part we will be using Retrofit2 and it’s Gson adapter so let’s start by adding their respective dependencies to our build.gradle file (also add the RxJava dependencies, as we’ll need them later):
implementation 'com.squareup.retrofit2:retrofit:2.4.0'
implementation 'com.squareup.retrofit2:converter-gson:2.4.0'implementation 'io.reactivex.rxjava2:rxandroid:2.0.2'
implementation 'io.reactivex.rxjava2:rxjava:2.1.13'
implementation 'com.squareup.retrofit2:adapter-rxjava2:2.4.0'
By the way, let’s not forget to add the Internet permission to our manifest file:
<uses-permission android:
Retrofit makes it really easy to consume APIs by doing a lot of heavy lifting for us. It won’t do all the work by itself though, so let’s see what we have to do to get things rolling:
- Configure and build Retrofit
We use builder pattern to configure and build a Retrofit instance. For more details about Retrofit configuration see the documentation, for now just make sure to include API’s base URL, GsonConverterFactory and RxJava2CallAdapterFactory. These are necessary if we want to use Retrofit with Gson and RxJava, as we will be doing in this example.
Retrofit retrofit = new Retrofit.Builder()
.baseUrl()
.addConverterFactory(GsonConverterFactory.create())
.addCallAdapterFactory(RxJava2CallAdapterFactory.create())
.build();
If you’re more familiar with some other JSON parsing library or just don’t feel like using Gson you can check out the list of supported converters here.
2. Create a model class
Gson uses model (or POJO) classes to convert JSON data to objects. One thing to remember when it comes to model classes is this: they have to reflect the structure of JSON response we want them to represent. So let’s say we want to get some data about a famous movie director from the popular and free-to-use The Movie Database API — if we consult the API documentation, we can see that our JSON response would look something like this (shortened for the sake of simplicity):
{
"id": 287,
"imdb_id": "nm0000040",
"name": "Stanley Kubrick",
"place_of_birth": "New York City - USA"
}
We could write the model class by hand but there are many convenient converter tools to save us some precious time, for example this one. Copy the response code, paste it into the converter, select Gson, click on preview and, if all went well, what comes out should very much resemble a simple POJO class:
public class Person { @Expose
@SerializedName("id")
private Integer id;
@Expose
@SerializedName("imdb_id")
private String imdbId;
@Expose
@SerializedName("name")
private String name;
@Expose
@SerializedName("place_of_birth")
private String placeOfBirth;
// bunch of boring getters and setters
}
As we can see, the fields represent our JSON response. Annotations are not necessary for Gson to work but they can come in handy if we want to be specific or keep the camel case naming. Just keep in mind that if a variable name doesn’t match the JSON key of the corresponding key/value pair Gson won’t recognize it and it will be ignored. That’s where the @SerializedName comes in — as long as we specify it correctly we can name the variable however we want. And that’s that for the model class. Let’s see how to actually make use of Retrofit to specify our requests.
3. Create an ApiService interface
To use Retrofit we simply make an interface to house our API requests (name of the interface is arbitrary, but it’s a good practice to end it with service). Then we use annotations from the Retrofit library to annotate interface methods with specifics of the each request. Going back to the API documentation, we find that our request should look something like this:
public interface ApiService { @GET("person/{person_id}")
Single<Person> getPersonData(@Path("person_id") int personId,
@Query("api_key") String apiKey);
}
Let’s take a closer look at this piece of code and see what it means:
- @GET tells Retrofit what kind of request this is, followed by the coresponding API endpoint. It should be obtained from the API documentation
- Single<Person> is an RxJava constuct and we’ll deal with it later, for now just keep in mind that it requires a type, which must be of the model class we are fetching with the request (in this case, Person)
- specific method name is not required, but we shouldn’t stray away from good naming conventions
- @Path and @Query annotations are used on method parameters to specify path and queries used to build the request
Put simply, when this method is called, Retrofit will generate a request and forward it to the API. If successful, it will use the Gson converter to convert the response to the object of our model class. Under the hood, the generated request looks roughly like this:{person_id}?api_key=<<api_key>>&language=en-US
For those still uncertain about what goes where, here is a quick recap:
- - this is the base URL
- person/{person_id} - this is the endpoint followed by a path
- ? - a question mark indicates the beginning of queries
- api_key=<<api_key>> - this is a query, in this case, for an api key
- & - more queries incoming
- language=en-US - this too is a query, but we left it at default
And since we annotated the parameters with @Path(“person_id”) and @Query(“api_key”), the arguments we pass when calling the method will replace {person_id} and <<api_key>>.
4. Hook up the service with Retrofit and make the request
All that’s left to do is to actually make the request. To do that we must first create an instance of the ApiService using the Retrofit object we created at the beginning.
// create an instance of the ApiService
ApiService apiService = retrofit.create(ApiService.class);// make a request by calling the corresponding method
Single<Person> person = apiService.getPersonData(personId, API_KEY);
That last line of code is where we actually make a request to get the data we need, wrapped in a Single object. But what is a Single? And how does it help us deal with the asynchronous nature of network calls? We answer these questions, and more, in the second part.
Part II: Handling asynchronicity and more with RxJava
We’ve all heard of AsyncTask. Most of us have used it at some point in our developer careers. The unfortunate few still use it today. But almost everybody hates it. It’s overly verbose, clumsy, prone to memory leaks and bugs, and yet, in lack of a better tool, many online tutorials still use it in dealing with all that background work Android platform is ripe with.
In this part we’ll see how to ditch the AsyncTask altogether in favor of RxJava, and in the process, save yourself a lot of time and avoid headache. Keep in mind that this is not an RxJava tutorial and we will only scratch the surface of what RxJava is capable of — working with Retrofit to ensure a simple and elegant solution to asynchronous network calls.
In the last part we left off at this line of code:
Single<Person> person = apiService.getPersonData(personId, API_KEY);
So what is a Single? Without making this an RxJava tutorial, let’s say it allows us to recieve a single set of data from the API, do some stuff with it in the background, and, when done, present it to the user — all that in a few lines of code. Internally, it is based on the observer pattern and some functional programming goodness, with data being pushed to interested observers at the moment of subscription.
To receive the data we only have to subscribe to a Single, calling the subscribe() method on it and passing a SingleObserver as an argument. SingleObserver is an interface containing 3 methods. By subscribing we ensure that the data is pushed when ready, and is passed to us in onSuccess() method. That is, if request was successfully completed — if not, onError() is invoked, enabling us to deal with the exception as we see fit.
Single<Person> person = apiService.getPersonData(personId, API_KEY)
.subscribe(new SingleObserver<Person>() {
@Override
public void onSubscribe(Disposable d) {
// we'll come back to this in a moment
}
@Override
public void onSuccess(Person person) {
// data is ready and we can update the UI
} @Override
public void onError(Throwable e) {
// oops, we best show some error message
}
});
But what about the onSubscribe() method? It is called in the moment of subscription and it can serve us to prevent potential memory leaks. It gives us access to a Disposable object, which is just a fancy name for the reference to the connection we established between our Single and a SingleObserver — the subscription. That subscription can be disposed with a simple method call, thus preventing those nasty situations when, for example, rotating the device in the middle of a running background task causes a memory leak. What we want to do is the following:
- first we create a CompositeDisposable object which acts as a container for disposables (think Recycle Bin) and add our Disposable to it in the onSubscribe() method:
@Override
public void onSubscribe(Disposable d) {
compositeDisposable.add(d);
}
- then we simply call the dispose() method on the CompositeDisposable in the appropriate lifecycle method:
@Override
protected void onDestroy() {
if (!compositeDisposable.isDisposed()) {
compositeDisposable.dispose();
}
super.onDestroy();
}
We’re almost there — we got the data we wanted and we did our chores of cleaning up any potential mess. But we still haven’t dealt with asynchronicity. NetworkingOnMainThread exception will still be thrown. Now comes the hard part, you must be thinking? Not really. Thanks to the RxAndroid library, RxJava can be made aware of Android threads, so all we have to do is add two more lines of code:
Single<Person> person = apiService.getPersonData(personId, API_KEY);
person.subscribeOn(Schedulers.io())
.observeOn(AndroidSchedulers.mainThread())
.subscribe(...);
As easy as that, with subscribeOn() we told RxJava to do all the work on the background(io) thread. When the work is done and our data is ready, observeOn() ensures that onSuccess() or onError() are called on the main thread.
While we’re at it, let’s explore RxJava some more. Let’s say, for the sake of example, that we want to make some more use of the API and fetch a list of movies related to our Person. We would have to make another API request following the same steps — create a model class and add another method to the ApiService, with the corresponding API endpoint. You’re probably thinking we have to repeat the steps with RxJava too, then think of a complex logic to deal with acychronicity and time the callbacks so we can update the UI at just the right time? Not necessarily.
Enter RxJava Operators. Operators are methods that allow us to, well, operate on our data and manipulate the way it’s being pushed to us. They can be called on Single objects and chained just like we did before. The list of available operators and their uses is huge and you can find it here.
In our case, wouldn’t it be nice if we could simply join the Person with the list of movies in the background and get them at the same time, as a single set of data? Luckily, there is just the Operator to meet our needs — Zip. Presuming the request for movies returns a list of Movie objects, we can add a List<Movie> field to the Person class and do the following:
Single<Person> person = apiService.getPersonData(personId, API_KEY);
person.zipWith(apiService.getMovies(personId),
(person, movies) -> {
person.setMovies(movies);
return person;
})
.subscribeOn(Schedulers.io())
.observeOn(AndroidSchedulers.mainThread()
.subscribe(...);
If it seems confusing at first glance, don’t worry, it usually is. It takes some time and practice to get acquainted with the syntax, and knowing your lambda expressions helps a lot. What we did is call the zipWith() method on the Single wrapping a Person request — zipWith() accepts 2 arguments — another Single (the one wrapping a list of movies) and a lambda expression which gives us access to both person and movie list data, since the Zip operator joins two separate data sources. In the lambda we assigned the movie list to the person. Now, when all the background work is done, onSuccess() will give us a Person object together with the list of movies we assigned to it.
Again, make sure to check out the list of operators, as there are many of them, and they can be used in a myriad of ways to achieve most wonderful things.
And that about covers it. We did what we set out to do, without going to much into the inner workings of any of the used libraries. Both Retrofit and RxJava are amazing tools and this article offers only a glimpse at the tip of the iceberg of what they’re capable of. Even so, if it managed to pique your interest and motivate you to dig below the surface we will consider the time spent writing it time well spent indeed.
If you want to get a better insight into how this stuff works in a simple MVP app check out my Github repo.
If you’re interested in the Room Persistence Library and how to make it work with RxJava, you can check out my next article on the subject:
SQLite on Android Made Simple: Room Persistence Library with a Touch of RxJava
Support for SQL relational databases has been built into the Android system since it’s early days, in the form of…
medium.com
Also, feel free to post any questions, comments, suggestions or corrections you may have, or even better — visit the official sites of RxJava and Retrofit and get more immersed into their amazing worlds. You’ll love it, and you’ll love yourself for doing so. | https://medium.com/android-news/rest-api-on-android-made-simple-or-how-i-learned-to-stop-worrying-and-love-the-rxjava-b3c2c949cad4?gi=f304035237d0 | CC-MAIN-2021-39 | refinedweb | 2,458 | 57.1 |
The pdb (for 'p'ython 'd'e'b'ugger) module enables you to step through executing code, pausing at each line, or running to pre-set breakpoints. Docs are at. If you're on Windows, make sure that you run
python and not
pythonw, or you won't see the output.
pdb is certainly low level, but after using it a bit you should be able to figure out how to get around pretty well. If you have a section of code you want to step through, add these two lines:
import pdb pdb.set_trace()
That's the equivalent to the VFP
SET STEP ON command, for those of you who are VisualFoxPro users. From that point, execution halts until you tell it how to proceed. The main keys and their Fox equivalents are:
- s - Step Into
- n - Step Over
- u - Step Out
- b NNN - sets a breakpoint at line
NNN
- c - Run (until the next breakpoint is reached)
Any time execution stops, you can issue Python commands to check the values of variables, properties, etc., or even change them. | http://dabodev.com/wiki/Pdb | crawl-001 | refinedweb | 180 | 77.57 |
Is This Content Helpful?
We're glad to know this article was helpful.
Instructions provided describe how to use the Python UpdateCursor method to calculate values in a table based on another field.
Sometimes it is desirable to set attributes in a field to certain values based on the value of another field. In this case the UpdateCursor with conditional statements can be used to correctly set those values.
Pseudo Code:
1. Create a geoprocessor object.
2. Input the feature class to be updated.
3. Create the UpdateCursor object.
4. Iterate through the rows in the table and update the values based on the conditional statements.
Code:
# Import the standard modules and create the geoprocessor...
import arcgisscripting, os, string, sys, time
gp = arcgisscripting.create()
# Set the overwrite to true...
gp.overwriteoutput = 1
# The file to be updated...
fc = r"C:\Temp\testdata.shp"
# Create the UpdateCursor Object
cur = gp.updatecursor(fc)
row = cur.next()
while row:
# Get the value of the field the calculation will be based on...
field1 = str(row.getvalue("FIELD1"))
# Conditional Statements
# if the field1 = value1, set field2 to equal NewVal1, etc...
if field1 == "Value1":
row.FIELD2 = "NewVal1"
cur.updaterow(row)
row = cur.next()
elif field1 == "Value2":
row.FIELD2 = "NewVal2"
cur.updaterow(row)
row = cur.next()
else:
row.FIELD2 = "NewVal3"
cur.updaterow(row)
row = cur.next()
del cur, row, gp | https://support.esri.com/en/technical-article/000010171 | CC-MAIN-2018-43 | refinedweb | 223 | 64.27 |
Fade text effect
In this tutorial you will learn how to create a fade text effect in Flash. The fade effect is basically text which fades in from being transparent. This effect is fairly basic and very simple to create, no actionscript is needed.
Fade text effect
Step 1
Open a new Flash document.
Select the Text tool with static text and type your message on the stage. You can choose whatever font type or font size you wish.
Step 2
Convert your text message in to a symbol by pressing F8. Then give your symbol an appropriate name, check movie clip and click ok.
Step 3
On the timeline insert a new key frame at frames 20 and 40 by selecting F6 at each frame. This creates a short pause after the text fade effect. You can make this pause shorter by decreasing the number of frames after the second key frame.
Using the selection tool (v) select the first frame. Then click your text message on the stage. Now change the colour to alpha at 0% at the bottom of the screen. This will make your movie clip transparent at the beginning of the animation.
Now create a motion tween in between the first and second key frames. On the timeline right click in between the first and second key frame and select “Create motion tween”. You time line should look like below:
Step 4
Test your movie clip Ctrl + enter.
You should now have a nice fade text effect. Checkout the video version of the fade text effect if you are having problems.
This fade text effect can also be achieved using Tweens with Actionscript code. The below code would be the alternative for step 3. For more information on Tweens checkout this tutorial.
import fl.transitions.Tween;
import fl.transitions.easing.*;
var myTween:Tween = new Tween(text_mc, "alpha", Strong.easeOut, 0, 1, 10, true);
12 comments:
Very clear and easy to understand. Thank you.
Hi again, I have a weird thing happening in flash.
After I created the text effects with your tutorial everything runs fine within Flash but when I publish it to HTML (or pressing CTRL+ENTER to view it in a Flash window) the effects are gone.
Have I done something wrong?
@Mau
Can you rephrase your question I’m not quite sure what you mean?
Sure, the thing is that the text effects works well under flash. I mean in the flash editor, when I press enter they work ok.
When I publish (pressing F12) or watch it (pressing CTRL+ENTER) the text effects are gone.
Would you like me to send you the flash file?
@Mau
Does this particular problem happen with other flash tutorials as well?
I don't know. I am really new to flash. I wanted to create a presentation this weekend, but I have been stuck with this issue for hours.
I don't think there's anything wrong with your tutorial either, I guess the problem is between the chair and the computer, I mean most probably the problem is myself.
Anyway thanks for your help. Have you had this problem before?
@Mau
I've never had this problem before.
Ok, I'm new a this, I did get the first line of text to fade, but I wanted to add a second line. I created a new layer, started the frame at 41, followed the fade instructions for the second line. When I perform a play back only the first line fades in and out. What am I doing wrong?
@Corliss
You need to create a empty key frame at frame 41. Then try the fade text.
I'm having a similar problem...Everything is set up correctly I believe, but when i export the clip it doesn't show the text fading. Within the editor itself it does show it fade.
@Nick
When you export the movie make sure you set the file type as Swf. If this doesn’t work try to publish the movie.
Happened to me too. When you are creating the text symbol make sure the text type is NOT dynamic, but rather static. It works fine now. | http://www.ilike2flash.com/2008/06/fade-text-effect.html | CC-MAIN-2013-48 | refinedweb | 698 | 84.57 |
Using Google Channel API I find that /_ah/channel/disconnected/ is
always called promptly while/_ah/channel/connected/ is not. Many
times I never get a connect call and then get a disconnect call for a
channel which the server has not been notified about!
(I saw
some people had the opposite problem that disconnect is delayed)
Given a "vanity url" for a channel (e.g. youtube.com/wheelockcollege,
which points to youtube.com/wheelockmarketing1), is there a way to retrieve
the channel details?
I'm using the new version 3 of the API,
with the latest Python package.
I am using inbound-channel-adapter of spring integration. I want to poll
under two different directories -one per file category- and parse the files
that are located there. The code that i use is:
<int:channel<file:inbound-channel-adapter id="fileInOne"
directory="myDirOne"
i had problem with application messaging, ...after some question and
answer, i went to use Remoting namespace, under System.Runtime
namespace...
it worked perfectly, but the matter is when
application terminate with exception...if server stop in not manner,
the channel i register will stay registered, ...
i dont have
much knowledge over remoting or other related ma
I've created a matrix from a ARGB_8888 bitmap like this:
Mat detectedFaceMat = Utils.bitmapToMat(bmp1);
The resultant matrix has 4 channels. I now need to compare it to 1
channel matrices. Not sure how to convert the 4 channel to 1 (I only need
grayscale).
I'm completely new to audio in .NET, so bear with me.
My
goal is to create a single wav file with two channels. The left channel
will consist of a voice message (stream generated using SpeechSynthesizer),
and the right channel needs to be a simple tone at a given single
frequency.
To complicate things a little more, I also need for
the right channel tone to "bookend" th
I am creating a web application using Google App Engine with the Channel
API. This is my first attempt at doing anything Ajax-like, and I suspect
there is a standard technique for solving my problem, but I've been unable
to find it.
My client program will post messages to the server.
If the message is not able to be sent, for example, if the channel is
closed, what happens to th
I'm really stumped on this one. I have an image that was [BGR2GRAY]'d
earlier in my code, and now I need to add colored circles and such to it.
Of course this can't be done in a 1 channel matrix, ans I can't seem to
turn the damned thing back into 3.
numpy.dstack() crashes
everything
GRAY2BGR does not exist in opencv2
cv.merge(src1, src2, src3, dst) has been
I am trying to change 3-channel image into 4-channel like this:
cv::VideoCapture video;video.open("sample.avi");cv::Mat
source;cv::Mat newSrc;int from_to = { 0,0, 1,1, 2,2, 3,3 };for ( int i = 0; i < 1000; i ++ ){ video >>
source; cv::mixChannels ( source, 2, newSrc, 1, from_to, 4 );}
Then I
Here's my question
On my Main Stage I added this code to
have a sound repeat, it works fine on this stage and loops well
var HP1sound:Sound = new HP_sound();var
HP_channel:SoundChannel = new SoundChannel();function
playSound():void{ HP_channel=HP1sound.play();
HP_channel.addEventListener(Event.SOUND_COMPLETE, onComplete);}
Sound
Channel
repeat
plays
other
channel | http://bighow.org/tags/CHANNEL/1 | CC-MAIN-2018-05 | refinedweb | 571 | 62.78 |
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Published byAvery Manfull Modified about 1 year ago
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Functional Programming Universitatea Politehnica Bucuresti Adina Magda Florea
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Lecture No. 10 & 11 Type declarations Data declarations Recursive types Arithmetic expressions Binary trees Type inference I/O
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Note The slides are from: Programming in Haskell, Graham Hutton, Cambridge University Press (January 15, 2007)
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Type inference Type signatures are not mandatory. Haskell has a type inference system Type inference in Haskell is decidable isL c = c == 'l' This function takes a character and sees if it is an 'l' character.
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Type inference isL c = c == 'l' The compiler derives the type for isL something like the following (==) :: a -> a -> Bool 'l' :: Char Replacing the second ''a'' in the signature for (==) with the type of 'l': (==) :: Char -> Char -> Bool isL :: Char -> Bool the return value from the call to (==) becomes the return value of isL function.
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Type inference isL is a function which takes a single argument. We discovered that this argument must be of type Char. Finally, we derived that we return a Bool. So, we can confidently say that isL has the type: isL :: Char -> Bool isL c = c == 'l'
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Reasons to use type signatures Documentation: the most prominent reason is that it makes your code easier to read. Debugging: if you annotate a function with a type, then make a typo in the body of the function, the compiler will tell you at compile-time that your function is wrong. Types prevent errors
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Reasons to use type signatures fiveOrSix :: Bool -> Int fiveOrSix True = 5 fiveOrSix False = 6 pairToInt :: (Bool, String) -> Int pairToInt x = fiveOrSix (fst x) The function fiveOrSix takes a Bool. When pairToInt receives its arguments, it knows, because of the type signature that the first element of the pair is a Bool. Extract this using fst and pass that into fiveOrSix This would work, because the type of the first element of the pair and the type of the argument to fiveOrSix are the same.
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I/O Performing input/output in a purely functional language like Haskell has long been a fundamental problem. How to implement operations like getChar which returns the latest character that the user has typed or putChar c which prints the character c on the screen? We somehow have to capture that getChar also performs the side effect of interacting with the user.
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I/O Haskell's I/O system is built around a mathematical foundation: the monad. Monads are a conceptual structure into which I/O happens to fit. I/O operations are seen as actions Actions are defined rather than invoked within the expression language of Haskell.
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I/O actions The invocation of actions takes place outside of the expression evaluation.
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I/O actions Every I/O action returns a value. In the type system, the return value is "tagged" with IO type, distinguishing actions from other values. The type of the function getChar is: getChar :: IO Char The IO Char indicates that getChar, when invoked, performs some action which returns a character.
39
I/O actions Actions which return no interesting values use the unit type, ( ). For example, the putChar function: putChar :: Char -> IO () takes a character as an argument but returns nothing useful. The unit type is similar to void in other languages.
40
do Actions are sequenced using the operator >>= (or `bind'). Instead of using this operator directly we can use the do notation The keyword do introduces a sequence of statements which are executed in order. A statement is: an action a pattern bound to the result of an action using <- a set of local definitions introduced using let.
41
do A simple program to read and then print a character: main :: IO () main = do c <- getChar putChar c
42
do We can invoke actions and examine their results using do, but how do we return a value from a sequence of actions? Consider the ready function that reads a character and returns True if the character was a `y' ready :: IO Bool ready = do c <- getChar c == 'y' – Not correct !!!
43
return This doesn't work because the second statement in the 'do' is just a boolean value, not an action. We need to take this boolean and create an action that does nothing but return the boolean as its result. The return function does just that: return :: a -> IO a ready :: IO Bool ready = do c <- getChar return (c == 'y')
44
Examples mypair :: IO (Char, Char) mypair = do x <- getChar y <- getChar return (x,y)
45
Examples Reads a string from the keyboard getline :: IO String getline = do x <- getChar if x == '\n' then return [ ] else do xs <- getLine return (x:xs)
46
Examples An action that prompts for a string to be entered and displays its length: strlen :: IO () strlen = do putStr "Enter a string: " xs <- getLine putStr "The string has " putStr (show (length xs)) putStrLn " characters"
47
Examples!"
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© 2016 SlidePlayer.com Inc. | http://slideplayer.com/slide/4157475/ | CC-MAIN-2016-50 | refinedweb | 863 | 55.37 |
WTPNewsletter 20070518
Contents
WTP Weekly What's Cooking?
Headline News!
- WTP 2.0 RC0 declared for Europa M7.
- EMF feature changes
- JEE5 EMF models are moved to the org.eclipse.jst.j2ee.core plugin from org.eclipse.jst.jee for dependency requirements. (No package or namespace changes.)
- WTP now uses javax.wsdl 1.4 from Orbit.
WTP 2.0
May 11, 2007 - May1
Focus Areas
WTP 1.5.5
No builds yet.
May 11, 2007 - May 18, 2007
- Plugin Version Changes
- [ No Versioning Information Yet]
References
- Visit the WTP 2.0 Ramp Down plan for updated information on WTP 2.0.
Back to What's Cooking Archive
Back to Web Tools Project Wiki Home | http://wiki.eclipse.org/WTPNewsletter_20070518 | CC-MAIN-2016-44 | refinedweb | 114 | 72.32 |
Can't evaluate functions while paused at a breakpoint
Hi,
Just started using the Plugin Debugger package and I'm excited about what it can do.
Unfortunately I'm encountering one fairly major snag so far, though -- when Winpdb stops at the "spdb.start() line" (which I've placed in my plugin), I'm not able to evaluate statements in Winpdb involving any kind of sublime function call.
For example, using the test plugin code at the end of this post, when Winpdb stops at spdb.start(), if I enter this into the Winpdb REPL:
eval view.file_name()
then Winpdb waits a few seconds and comes back with:
The operation will continue to run in the background.
I presume this is because, while the plugin is paused, Sublime Text is also effectively blocked and can’t respond to anything?
If I allow Winpdb to continue execution, it will then evaluate these prior statements I’d entered and print results to the REPL. However, I feel this defeats the purpose a bit -- to effectively debug, ideally I'd see eval results while paused at the breakpoint.
Just wondering if anyone else has encountered a similar issue and, if so, if you might have discovered a workaround?
Using ST3 build 3142, Plugin Debugger 2014.04.19.16.30.06, Winpdb 1.4.8 (connected to an Anaconda Python 2.7 environment), Windows 10.
Here's the test plugin for the above example:
import os, sys import sublime, sublime_plugin class debugTestCommand(sublime_plugin.WindowCommand): # Constructor def __init__(self, window): sublime_plugin.WindowCommand.__init__(self, window) def run(self): view = sublime.active_window().active_view() import spdb ; spdb.start() print(view.file_name())
Many thanks in advance for the help!
Neil
Hi Neil,
I have tried to reproduce your issue and I found the reason for this:
What you can do is:
view.file_name() runs a sublime-text API command, but ST is frozen while the debugger is in break mode and will not respond. You can only eval() on python variables and functions.
You can use normal console for evaluating such expressions or you do it (as in example above) using a extra variable to store the result of a ST API function.
Good luck with your debugging session :)
Best regards,
Kiwi
Am 06.09.2017 um 01:39 schrieb Neil Alexander McQuarrie:
Hi Kiwi,
Thank you so much for the reply and for the idea -- I really appreciate it.
If I understand you correctly, if I want to make any kind of ST API call from within the debugger, I'll need to have already performed the call and assigned the result to a variable? Hmmm... in most of my own use cases I find that, when I'm debugging, it's an iterative process and I won't know going in what I'll be wanting to evaluate... (Or if I do, I might as well just put the statement into a print command and not bother with a debugger?)
Just confirming that I understand. Again, many thanks for your help.
Neil
I wonder if there's a programmatic way to have winpdb temporarily "un-freeze" / release control of ST without also advancing to the next line of code. I'm willing to research this and submit a PR if I can come up with something workable. No chance you'd have any leads / thoughts yourself as to whether this is a viable road to try to go down? | https://bitbucket.org/klorenz/plugindebugger/issues/8/cant-evaluate-functions-while-paused-at-a | CC-MAIN-2018-09 | refinedweb | 573 | 63.7 |
thanks!
I like the scrolling tab, it looks really clean (this was the main issue I had with the UI of SublimeText), but there is one small annoyance: if the tab are all "collapsed" on the left, it is easy to quickly scroll by alwasy cliking at the same point. But if they are all collapsed to the right, then the first one i need to click on a small region, and for the second one this now reveals the close button of a tab, and i need to move the mouse to scroll more. One thing I would like is the ability to scroll the tab using the scrollwheel when the mouse is hovering the tabs (should be pretty easy to add i guess), then it would really be perfect
Edit: I noticed that you can keep the button clicked on the "collapsed" tab and this will make the different scroll. Not as good as a scrollwheel, but at least usable when i want to scroll from right.
That would be pretty pretty cool!
Also good to see that a window now closes if the last tab is dragged out.
If you add mouse wheel support, you can probably remove those left/right arrow buttons.They take up quite a lot of space but are not that useful in their current form as it takesmany clicks to scroll to the tab that's wanted.
The dropdown is nice. Maybe hide it if all tabs are fully visible?
Reading your comment I disabled the Soda Theme i was using and i noticed the arrow, and the drop down, that do not exist with Soda.Drop down is nice indeed.An option to disabled the arrow would be good, and I like the auto-hide idea for the drop-down / arrow if not needed.
Edit: Oh, after a disable/enable of the theme, arrow and drop down are available in Soda
Same here, Soda removes it.But with default theme, clicking on Left/Right arrows seems to do nothing.
Re-enabling Soda makes arrows still visible until you restart ST3, this is more of a bug I guess. (Hot theme switching is kind of buggy for me)
EDIT: Oh and the option about the minimap, I did not find it in the default conf file!
If you add minimap_scroll_to_clicked_text to the default Preferences you might as well addwide_caret while at it. It's very useful and still works...
These 2 issues are still present:
jps, could you please give me an answer on the first one ?
The behavior I'm getting (with the Flatland Dark theme, Windows 7 x64) is that if I turn the theme off, then back on, I get the arrows and dropdown. If I then restart Sublime, I lose the arrows and dropdown. That being said, the dropdown works but the arrows do not. When I add additional new tabs they continue to shrink instead of scrolling. Not a critical issue for me as I rarely have enough tabs open to scroll through them, but thought I'd add in my experience.
Loving the tab scrolling. Works really well on OSX with both magic mouse and track pad.
I'm noticing random text sizes changes (maybe once an hour), as if I was holding down ctrl and moving the scroll wheel, even though I am not. Did not notice it with <3048. Additionally, it seems the scroll wheel on the mouse has gotten super-touchy - when I scroll down what used to move a few lines, it now scrolls hundreds of lines.
I am running the x64 tarball on Arch with Gnome Shell.
With this build, on OSX, when I save a root file it will just notify me that it's read only instead of asking for my password like in previous builds.
tabscrolling is cool, but it should also work with mousewheel scrolling if you hover over the arrows or better over the whole tabset area
Thanks so much for the minimap_scroll_to_clicked_text setting!
Clicking on the arrows doesn't do anything on Windows 7.
What is the expected behavior with this setting enabled? Can't find any difference.
There is a race condition in handling Settings.add_on_change callbacks. If you call settings.clear_on_change in the callback with the current key, Sublime dies with an access violation.
Code to reproduce:
[code]class AutoFormatSettings(sublime_plugin.EventListener):
event_key = 'auto_format_settings'
def on_post_save(self, view):
name = view.file_name()
if not name:
return
base = os.path.basename(name)
_, ext = os.path.splitext(base)
if ext == '.sublime-settings':
settings = sublime.load_settings(base)
callback = functools.partial(self.on_reload, base)
settings.add_on_change(self.event_key, callback)
def on_reload(self, base):
settings = sublime.load_settings(base)
settings.clear_on_change(self.event_key) # <<< here ST dies
sublime.save_settings(base)[/code]
I have noticed a bug in this build, on Windows 7 when "atomic_save" is set to false, saved files (new or changed) become hidden. (File attribute "hidden" is set)
This behaviour is introduced in 3048, the previous build saves the files normal (not hidden).
With this build (or possible earlier, as I hadn't updated for some time) ST3 stops rendering if I started it when the integrated gfx chip was active on my macbook and it then changes to the discrete. Changing back to the integrated makes it render again.
If I instead start ST3 with the discrete chip enabled and it switches to integrated, it's still rendering but is now very choppy. It's back to full speed again if the application is restarted or I switch back to discrete.
Reproducible via the gfxStatus app.
Annoying. I've set so many files to read-only lately due to that. While read-only attribute doesn't bother Sublime (it still is able to save), all other editors (VS) complain that they can't save the file. I have to manually remove hidden attribute.
+1 to this, still happen! | https://forum.sublimetext.com/t/dev-build-3048/10685/15 | CC-MAIN-2016-44 | refinedweb | 981 | 72.76 |
Capture the Image Using Ultrasonic Sensor With Arduino
Introduction: Capture the Image Using Ultrasonic Sensor With Arduino
I have been in IOT space for quite few months and trying to integrate things with Arduino board, Recently I came across Ultrasonic sensor, it is interesting. So I thought of creating a small project.The project goal is to capture the obstacle for security purpose using ultrasonic sensor with a camera.
Hardware used :
1) Ultrasonic sensor
2) Arduino UNO
3) Camera
Step 1: Hardware Explanation
Ultrasonic sensor
Ultrasonic sensor converts sound wave into electrical signal, they do both transmitting and receiving the signal, It will act like as an Transducer.Ultrasonic generates high frequency sound waves so the echo is received back to the sensor in between the transmit time and receiving time is calculated by the arduino and it will give the input to python. For reference you need more information about ultrasonic visit this link Ultrasonic.
Arduino UNO
Arduino is an open-source prototyping platform based on easy-to-use hardware and software.There are different types of arduino board available as per our requirements we select the boards in this project i am choosing Arduino UNO.
Camera
In this project i am using web camera that capture the image based on the detection.
Step 2: Arduino Code
Arduino will receive the signal from Ultrasonic and given the signal input to python.
int trigger_pin = 13;
int echo_pin = 11;
float time_taken;
void setup() {
Serial.begin(9600);
pinMode(trigger_pin, OUTPUT);
pinMode(echo_pin, INPUT);
}
void loop() {
digitalWrite(trigger_pin, LOW);
delayMicroseconds(2000);
digitalWrite(trigger_pin, HIGH);
delayMicroseconds(10);
digitalWrite(trigger_pin, LOW);
time_taken = pulseIn(echo_pin, HIGH);
Serial.println(time_taken);
delay(50);
}
Step 3: Python Code
Python program is used for getting the input signal from sensor via arduino, so that it can capture the obstacle according to the sensor detection.
#! /usr/bin/env python
import sys
import serial
import pygame
import pygame.camera
from os import getenv
from pygame.locals import *
from datetime import datetime as dt
# Initializing the Camera device
pygame.camera.init()
cam = pygame.camera.Camera("/dev/video0", (640, 480)) // Here declare the arduino port
home_dir = getenv('HOME')
'''
Adjust the value of this variable to set the distance for the sensor to detect intruders
'''
RANGE = 300
def capture_image(): ''' Starts the camera, Captures the image, saves it &amp;amp;amp; stops '''
file_name = home_dir + '/image_captured/image_' + str(dt.now()) + '.jpg'
cam.start() image = cam.get_image()
pygame.image.save(image, file_name)
cam.stop()
'''
Establishes a connection to Arduino board through serial interface
'''
arduino_board = serial.Serial(sys.argv[1], 9600)
'''
Enters an infite loop that runs until it receives Keyboard Interrupt
'''
while True:
if arduino_board.inWaiting() > 0:
data = arduino_board.readline().strip()
try:
'''
The value received through serial interface would be string, in order to process futher, it is converted to numeric datatype.
'''
data = int(float(data))
if data <= RANGE:
capture_image()
print data
except BaseException, be:
'''
initially the board might send some strings that are not the numeric value, to handle such exception it is catched and ignored by printing an exception message.
'''
print be.message
Step 4: Run the Program
Declare the arduino port in Python program the above image shows the arduino UNO port connection.
For running the program save the python code, open terminal type => python “Your python project name”/arduino port name (example : python self.py /dev/ttys0 ). Arduino port name is shown in arduino ide choose Tools => Port => Port name is shown in ide.
Once all these settings are done, When you run the program Ultrasonic sensor will find the obstacles in an interval and capture the images using the camera.Hope this will give you some idea about using ultrasonic sensor with arduino using Python.
is it possible if ultrasonik sensor changes to PIR sensor?
where can see result capture of image? do you use serial com to connect arduino to PC?
Hi
I don't try using PIR sensor but it is possible.
Capture image save in the home/image_captured
Yes i use serial port for communication
hi..
may i know your email?
i need your help about image processing..
karthikeyan.s@spritle.com
hi can you please share your present email id?
If I understand correctly this can be adjusted to accommodate different trigger distances? I ask because I am looking to create an image capture system for a 3D printer after the completion of every build layer. And I think this might be the starting point for me.
Great work! Would you happen to have an example image using this technique that you can share?
I Just have sample Images
plzz send a code siir
tamilarasank96@gmail.com
Please check your email
Cool project. | http://www.instructables.com/id/Using-ultrasonic-sensor-to-capture-the-image/ | CC-MAIN-2017-51 | refinedweb | 775 | 55.03 |
Pull requests
Implement periodic kernel (isotropic and 94c978de4652 hg commit -m 'Merged in kiudee/bayesopt (pull request #7)'
Thanks for the contribution!
Looking at the literature, the most common way of periodic kernel is equation (4.31) of Rassmusen and Williams That's a totally different kernel. This one is stationary (although your implementation miss the norm part) and RW is non-stationary.
Just a couple of details of implementation: - The order of parameters should be consistent. For 1D, ARD and ISO should be equivalent. For consistency with other kernels, the scale should go always first. - The parameter should be l, not l^2. - This is a header file. Put the constant pi in a local scope (e.g.: inside the operator) and never outside the namespace. - I haven't found a periodic kernel with ARD in any paper, but having a different "p" and "l" per dimension can be problematic. RW version is trivial to implement as ARD because it's a just modified SE kernel. | https://bitbucket.org/rmcantin/bayesopt/pull-requests/7/implement-periodic-kernel-isotropic-and/commits | CC-MAIN-2017-34 | refinedweb | 167 | 59.09 |
javax.xml.namespace.QName; 23 24 /** 25 * Class that defines groups at the schema level that are referenced 26 * from the complex types. Groups a set of element declarations so that 27 * they can be incorporated as a group into complex type definitions. 28 * Represents the World Wide Web Consortium (W3C) group element. 29 */ 30 31 public class XmlSchemaGroup extends XmlSchemaAnnotated { 32 33 /** 34 * Creates new XmlSchemaGroup 35 */ 36 public XmlSchemaGroup() { 37 } 38 39 QName name; 40 XmlSchemaGroupBase particle; 41 42 public QName getName() { 43 return name; 44 } 45 46 public void setName(QName name) { 47 this.name = name; 48 } 49 50 public XmlSchemaGroupBase getParticle() { 51 return particle; 52 } 53 54 public void setParticle(XmlSchemaGroupBase particle) { 55 this.particle = particle; 56 } 57 58 } | http://ws.apache.org/commons/xmlschema14/xref/org/apache/ws/commons/schema/XmlSchemaGroup.html | CC-MAIN-2015-27 | refinedweb | 123 | 51.89 |
Pattern Matching In Elixir
Pattern Matching in Elixir - Does it fit?
Pattern matching is a very powerful feature in any programming language that implements it, I think that essentially because pattern matching is inherent in computer programming logic, from a conceptual to a practical level. It's a feature that is present in many functional languages for the programmer to use. Elixir (inheriting from Erlang) gives us access to it and idiomatic code is most often peppered with pattern matching.
It fits well in many places and allows us to describe very succinctly the code paths our running programs may take, while retaining, and most of the time increasing, the clarity of the code. While some features that allow succinctness and conciseness of code end up in a symbolical soup more close to mathematics than a "legible" declaration (for those of us not completely familiar with mathematical notation), pattern matching does not suffer from that, because what it allows us to express is the "form" of the data inside our running programs, which in turn allows us to express what pathways of code should be run given that form.
Erlang itself has on top of that a very accessible and intuitive way of treating binary data, that coupled with pattern matching gives it the ability to chop, split, and work with binaries in a very pleasant way.
Now I could go on talking about how the rainbow is perfect during this sunset, or how unicorns are pooping diamonds and so on, but that won't give you any insight into Pattern Matching. So let's move to do some exploration of it.
First we'll start by analysing the
= operator in Elixir, which is in fact a match operator and not an assignment operator (if you're coming from Javascript, Ruby, etc). It has one side effect that makes it look like an assignment operator though, which is, when the expression on the left side of it is a
variable, it binds whatever is on the right side to that variable.
We can do some simple tests of it on iex:
1 = 1 # 1 1 = 2 # ** (MatchError) no match of right hand side value: 2 a = 1 # 1 a = 2 # 2 a + 1 # 3 a # 2 3 = a #** (MatchError) no match of right hand side value: 2 a = 3 # 3 a = a + 1 # 4 a # 4
Ok, so what can we see from these examples? First we see that when a match evaluates correctly, we get the value of the expression that was matched.
1 = 1 returns 1, and not true. If we wanted to check for equality we would use the equality operator
== and not the match operator
=.
> 1 == 1 # true > 1 == 2 # false
On the second case we get a
MatchError, because we're trying to convince our program that
1 matches
2, and it simply states, dude, perhaps in your universe, but in mine, 1 can't match 2.
Then we do a bunch of "assignments", and it works mostly as one would expect in other languages, the variable
a gets bound to whatever value.
When we try to do
3 = a, when
a was bound to 2, we get again an error, but when we switch it to
a = 3 we no longer get an error, instead the value 3 is bound to the variable
a.
In Erlang you can't use the same variable name in different assignments, but Elixir chose to allow it. Underneath, because the data is still immutable, what happens is that elixir creates various versions of
a and points
a to the last version you entered - from a practical point of view though, it looks like you're "re-assigning" the variable.
Now you might be thinking, all of that for what? Yes, it's not apparent yet why it helps in anything else other than silly examples. So let's move on.
Let's create a map and play around with it.
a = %{key_one: "this is key number one", key_two: "this is key number two"} # %{key_one: "this is key number one", key_two: "this is key number two"} %{key_one: key_one_var} = a # %{key_one: "this is key number one", key_two: "this is key number two"} key_one_var # "this is key number one"
So here things start to look more interesting. You can see that we placed a pattern of
%{key_one: key_one_var} and matched it against the previously map bound to
a. The match succeeded, but we also "bound" the variable
key_one_var to whatever was the value of the key
:key_one used in the pattern match.
If you used Javascript E6 destructuring you might notice it looks familiar. The Erlang (&Elixir) version though is way more powerful than javascript's version, because of its properties and how it can be used, but lets move on.
%{key_three: key_three_var} = a # ** (MatchError) no match of right hand side value: %{key_one: "this is key number one", key_two: "this is key number two"}
Ok, we tried to match the map bound to
a to a map of the form
%{key_three: some_variable} and this didn't work, because
a doesn't have any key named
:key_three, it's expected to fail.
But notice that we didn't need to specify all the keys when matching
:key_one previously, although the
a map had an additional
:key_two.
So how can we look at it conceptually? I think that the best way to describe it is, given
a = b, could the form described in
a be extracted from the contents held in
b? If it can the match succeeds, any bindings are made effective, and the full expression is returned. If it can't, an error is thrown.
And in this case indeed it can, in the previous example, the form
map, with a key named :key_one, to be bound to a variable named key_one_var can be extracted from
a, because that map has a key named
:key_one and we aren't specifying any value that that key must have.
So if we write:
%{key_one: "this is key number one"} = a
It also matches. But if we write:
%{key_one: "this is key number two"} = a ** (MatchError) no match of right hand side value: %{key_one: "this is key number one", key_two: "this is key number two"}
It fails, because we're saying match a map that has a key named
:key_one with the value
"this is key number two", but since the value of that key in
a is actually
"this is key number one" it fails.
%{key_one: "this is key number " <> keynumber} = a # %{key_one: "this is key number one", key_two: "this is key number two"} keynumber # "one"
Here it matches, and not only that, we have extracted the last portion of the binary string as a variable. We could extract both the last portion and the whole binary string if we wanted with for instance:
%{key_one: "this is key number " <> keynumber = whole_binary} = a # %{key_one: "this is key number one", key_two: "this is key number two"} keynumber # "one" whole_binary # "this is key number one"
It doesn't end here though, because the patterns you use can be much more complex and involve nested levels of maps, binary specifications, and lists, and so on, you can "describe" in as much detail as you want their forms and also extract very easily what you need from them.
%{} = a
Matches alright again, because we're simply saying, is the value bound to
a a map? And since it is, all is good. You might have thought that it would fail, because
%{} could be interpreted as an
empty map but maps behave in this regard a bit differently. To assert that a map is indeed empty, you need to compare it,
%{} == a would return
false for instance.
Now lists:
b = ["a", "list", :of, :stuff] # ["a", "list", :of, :stuff] [head | tail] = b # ["a", "list", :of, :stuff] head # "a" tail # ["list", :of, :stuff]
Here we bound
b to a list of 4 elements. Then we pattern-matched
b against the pattern
[head | tail]. Lists in Erlang (and consequently in Elixir) are like LISP lists, they're composed of cons cells, that can be thought of as a group of cells, where each cell holds a value and also a "pointer" to the next cell in the list.
In this case we can think of it as a structure where the first "cell" is
"a", which points to the cell with value
"list", which in turn points to the cell
:of, which points to cell
:stuff which in turn points to the end of the list (an empty list,
[]).
So a list is a collection of cells where each element holds its value and points to the next element, that's why they're called
cons cells and usually described as
(x . points_to_y)(y . points_to_z)(z . points_to_empty_list)(), or more correctly
(x . (y . (z . ()))). In Elixir and Erlang it looks like
[x | [y | [z | []]]].
So when we match
b to the pattern
[head | tail] what we're actually saying is, does a pattern of a non empty list (meaning where there is at least one cell, the
head) match
b? If it does, bind the value of the first cell of that list to the variable
head and the remaining of the list to the variable
tail.
If we try:
c = [] # [] [head | tail] = c # ** (MatchError) no match of right hand side value: []
We get a match error, in this case we're trying to match a non-empty list (the pattern we wrote specifies it should have one element at least,
head, pointing to a tail), against an empty list (
c) so we get a match error, because an empty list has no cells, it's itself the end of the list. This contrasts with the previous seen case of the map, where the empty map
%{} still matched ok, but in practical (and theoretical) terms it makes sense, once you start using it you'll see that the difference between what meaning an empty list and an empty map usually assume and the ability to match on simply being a
map warrant this (seemingly) small contradiction. Also, if a match with
%{} equated to saying an empty map, then when you were matching non-empty maps you would need to spell out all the keys in the map and that effectively and that completely defeats the pragmatic purpose of matches - or you would need to specify different behaviour for when using the pattern
%{}.
d = ["non_empty"] # ["non_empty"] [head | tail] = d # ["non_empty"] head # "non_empty" tail # []
In this case though, we bound a list with one cell,
"non_empty" to
d and when we pattern matched, it worked.
"non_empty" got bound to
head and the end of the list (an empty list) got bound to
tail.
Again, it's not super impressive (yet) but we'll get there soon. Let's see one more example before though:
deep = %{list_key: [:a_list, %{super_deep: [:a]}], date: "2019-05-01"} # %{date: "2019-05-01", list_key: [:a_list, %{super_deep: [:a]}]} %{date: <<year::binary-size(4), "-", month::binary-size(2), "-", day::binary-size(2)>>, list_key: [_, %{super_deep: [first_element_of_super_deep | t]}]} # %{date: "2019-05-01", list_key: [:a_list, %{super_deep: [:a]}]} > year # "2019" > month # "05" > day # "01" > first_element_of_super_deep # :a
Wow. So with a simple pattern matching, we were able to extract a lot of information as you can see. We didn't need to split a string to get all the pieces of the date, we didn't need to iterate on the list to get it's nested elements and we plucked an element from inside a list, that was inside a map, that was inside another list itself inside another map. We can also write it in a way that is more readable
%{ date: << year::binary-size(4), "-", month::binary-size(2), "-", day::binary-size(2) >>, list_key: [ _, %{ this_will_fail: [ first_element_of_super_deep | t ] } ] } = deep # ** (MatchError) no match of right hand side value: %{date: "2019-05-01", list_key: [:a_list, %{super_deep: [:a]}]}
Here, in the pattern we described, we changed the key name inside the map inside the nested list to
this_will_fail and it no longer matched, although everything else was the same as before.
So, this is a bit more impressive, although, still doesn't look very useful if we can only use this on match operations. Where it becomes really, really, useful is when we use it in conjunction with Elixir's (Erlang) ability to have multiple function definitions and/or inside some constructs the language provides, such as
case statements (Well actually, most places, to be sincere, allow their usage).
You might also have noticed that we used
_ in these last pattern matches.
_ (or any variable name starting with
_) tells the compiler we're not interested in that value, so it won't bind that variable, although it still requires something to be there.
So let's look at patterns with
case statements, since they're very common too.
a = [:a, :list] # [:a, :list] case a do [:a, :list] -> "a list with :a and :list" _ -> "something else" end # "a list with :a and :list" a = [:b, :list] # [:b, :list] case a do [:a, :list] -> "a list with :a and :list" _ -> "something else" end # "something else" case a do [_, :list] -> "a two element list where the second element is :list" _ -> "something else" end # "a two element list where the second element is :list" case a do [_, :list | _t] -> "a list with at least two elements where the second element is :list" _ -> "something else" end # "a list with at least two elements where the second element is :list" case a do [_] -> "a list with a single element" [_, _] -> "a list with 2 elements" _ -> "something else" end # "a list with 2 elements" case a do [_] -> "a list with a single element" [:b, :something_else] -> "a list with 2 elements, :b and :something else" [:b, :list | tail] -> "a list with 2 elements and a tail" end # "a list with 2 elements and a tail" tail # []
Since lists are made of cons cells, this matched on the last branch, because although we only had 2 elements in the list it technically is 3, since a proper list always ends in an empty list itself. So by having a list with 2 elements what we actually have is:
(element_1 . points_to_element_2)(element_2 . points_to_end_of_list)() <- this last element is itself an empty list.
If we did this, on the other hand:
case a do [_] -> "a list with a single element" [:b, :something_else] -> "a list with 2 elements" [:b, :list, _last_element] -> "a list with 3 elements" end
We get instead:
** (CaseClauseError) no case clause matching: [:b, :list]
So although the last element is present, given that it is the list termination (empty list) it doesn't count as an "actual" element, but just as the termination of the list. Notice that instead of separating the last element with
| like we did previously, we separated it with a
, (comma), effectively indicating our pattern required 3 actual elements, and not 2 elements and a tail.
Because the empty list that signals the end of a list is not itself an "element" it didn't match, but when we use
| it does match, because the tail might be a cons cell or the end of the list itself - the
| separator means the "next pointer" of the list, while
, means an actual element.
We also see that because no branch of the case could match, that we got a
CaseClauseError, which just means, given the expression passed on to
case, I couldn't find any "conforming" branch in those you've defined.
As you see the logic is the same as with the match operator we've seen before, but instead of getting
MatchErrors when it doesn't match we get
CaseClauseErrors.
What happens when we use
case is that the expression in
case EXPRESSION do gets matched against each branch of the case statement, so it's translated into something similar to (for illustration purposes):
# when `a` is [:b, :list] > case a do >> [_] = [:b, :list] -> .... >> [:b, :something_else] = [:b, :list] -> ... >> [:b, :list | tail] = [:b, :list] -> ... > end
So the first branch
[_] (list with a single element) doesn't match, neither does the second branch
[:b, :something_else], but the third does, because
a is effectively in the form
[:b | [:list | []]]. Or
[:b, :list | []]
We can see some more examples, now with binary matching.
date = "01-02-2019" # "01-02-2019" case date do "01-02-" <> year -> "it's year #{year}" _ -> "not sure" end # "it's year 2019" case date do <<day::binary-size(3), month::binary-size(3), year::binary-size(4)>> -> "it's day #{day} in month #{month} and year #{year}" _ -> "don't know" end # "it's day 01- in month 02- and year 2019" case date do <<day::binary-size(2), "-", month::binary-size(2), "-", year::binary-size(4)>> -> "it's day #{day} in month #{month} and year #{year}" _ -> "don't know" end # "it's day 01 in month 02 and year 2019" date = "01/02/2019" # "01/02/2019" case date do <<day::binary-size(2), "-", month::binary-size(2), "-", year::binary-size(4)>> -> "it's day #{day} in month #{month} and year #{year}" _ -> "don't know" end # "don't know"
So this is way more interesting, because now we can start seeing ways to drive the "flow" of our programs by defining the form the data should have.
When we look at functions, all the same concepts apply, but instead of being inside the body of a function, they're used to define what "branch" of the function should be "called" when passed arguments of a certain form.
Let's create a file somewhere. The following examples will all need to be placed inside that module but we'll omit it in the examples, and then the whole module must be copied to your iex shell before running the examples.
# some_file.ex defmodule PatternMatching do def test_1(:a), do: "Function matched on :a" def test_1([]), do: "Function matched on empty list" def test_1([element]) do IO.puts("Function matched on list with a single element: #{inspect element}") test_1(element) end def test_1([head| t]) do IO.puts("Function matched on non_empty list, with head: #{inspect head} and tail #{inspect tail}") test_1(t) end def test_1(element), do: "Function with one non-list argument: #{inspect element}" end
Copy that into the iex shell and you should see something like:
# {:module, PatternMatching, <<70, 79, 82, 49, 0, 0, 7, 128, 66, 69, 65, 77, 65, 116, 85, 56, 0, 0, 0, 189, 0, 0, 0, 20, 22, 69, 108, 105, 120, 105, 114, 46, 80, 97, 116, 116, 101, 114, 110, 77, 97, 116, 99, 104, 105, 110, 103, ...>>, {:test_1, 1}}
Now let's try calling some functions:
PatternMatching.test_1(:a) # "Function matched on :a" PatternMatching.test_1(:b) # "Function with one non-list argument: :b" PatternMatching.test_1([:b]) # Function matched on list with a single element: :b # "Function with one non-list argument: :b" PatternMatching.test_1([:a]) # Function matched on list with a single element: :a # "Function matched on :a" PatternMatching.test_1([:b, :b]) # Function matched on non_empty list, with head: :b and tail [:b] # Function matched on list with a single element: :b # "Function with one non-list argument: :b"
Now if we move
def test_1(element), do: "Function with one non-list argument: #{inspect element}"
to be the first function defined (and then copy the module again into iex) and run the same function calls we did before:
PatternMatching.test_1(:a) # "Function with one non-list argument: :a" PatternMatching.test_1([:b, :a]) # "Function with one non-list argument: [:b, :a]"
Now everything matches the first one, because it has no defined pattern it accepts everything so none of the other functions we defined get the chance to be tested and consequently run.
If we add this one function to the end:
def test_1(argument_1, argument_2), do: "Matched with 2 arguments: 1: #{inspect argument_1} #### 2: #{inspect argument_2}"
`
Then copy the module again to iex and run:
PatternMatching.test_1("arg1", ["arg", "2"]) # "Matched with 2 arguments: 1: \"arg1\" #### 2: [\"arg\", \"2\"]"
Now because we're passing two actual arguments, none of the others will match (since they have arity of 1, meaning the number of arguments they accept is 1) and since only this last one accepts 2 arguments, only this one will match.
If we do:
PatternMatching.test_1("arg1", ["arg", "2"], "arg3") # ** (UndefinedFunctionError) function PatternMatching.test_1/3 is undefined or private. Did you mean one of: # * test_1/1 # * test_1/2 # PatternMatching.test_1("arg1", ["arg", "2"], "arg3")
This basically means, "I'm sorry, I couldn't find a function test_1 with arity 3 in the module PatternMatching". It also hints, if it can find suitable hypotheses, other functions that are available in that module. In this case it can see that we have a function named "test_1" with both arity 1 and 2, so perhaps we might have tried to call that but got the arity wrong.
So by now we have seen some use cases for pattern matching and we've learned a bit about them, they will match the first case that has a conforming pattern, be it a branch of a
case statement or a function definition.
This means, that the order in which we define the branches or functions has a meaning as well. And also that they will raise specific errors when it can't find an actionable code path.
If we define the branch or function as simply a bound (or unbound,
_, variable) then it will match everything.
So our patterns must go from the most explicit to the least explicit in order to be unambiguous and actually describe the flow of our program.
In
case statements, the whole expression is matched as a single element, while in functions there's also arity to take into account.
One last piece of functionality that we can use in pattern matching is guard clauses. These allow us to extend further our pattern matching capabilities. These guard clauses allow the usage of a subset of kernel functions that are "special", in the sense that they are pure functions and guaranteed to be "fast". So for instance we might want to discern if the value is a list or a map, if it's a map we also want to discern between an empty map and a non-empty map, but in the case of it being a list we don't care if it's a list with 0 or more elements, just that it's a list. We could write it as such:
a_list = [] case a_list do expression when is_list(expression) -> "it's a list" %{} = expression when expression == %{} -> "it's an empty map" %{} = expression -> "it's a non-empty map" end # "it's a list" a_list = [:two, :elements] case a_list do expression when is_list(expression) -> "it's a list" %{} = expression when expression == %{} -> "it's an empty map" %{} = expression -> "it's a non-empty map" end # "it's a list" a_list = %{} case a_list do expression when is_list(expression) -> "it's a list" %{} = expression when expression == %{} -> "it's an empty map" %{} = expression -> "it's a non-empty map" end # "it's an empty map" a_list = %{some_key: 1} case a_list do expression when is_list(expression) -> "it's a list" %{} = expression when expression == %{} -> "it's an empty map" %{} = expression -> "it's a non-empty map" end # "it's a non-empty map" If we switch the last two branches and set `a_list` variable to an empty map we get: a_list = %{} case a_list do expression when is_list(expression) -> "it's a list" %{} = expression -> "it's a non-empty map #{inspect expression}" %{} = expression when expression == %{} -> "it's an empty map" end # "it's a non-empty map %{}"
Although the map is empty, because the 2nd branch matches any map, empty or not, it's that one that gets evaluated when the expression is a map, and the 3rd branch now has no chance to be tested.
So now let's see some real cases where we can use this.
Let's define a module and structure that holds users and the type of users they are along with their age. We'll also define some functions to work with lists of users.
defmodule User do defstruct [:name, :age, type: :regular] def count_older_than(list, age) do count_older_than(list, age, 0) end def count_older_than([], _, count), do: count def count_older_than( [%User{age: user_age} = user | t], age, count ) when user_age > age do count_older_than(t, age, count + 1) end def count_older_than([_ | t], age, acc), do: count_older_than(t, age, acc) end
And now let's create some users and a list of them:
user_1 = %User{name: "John", age: 25} user_2 = %User{name: "Doris", age: 30, type: :administrator} user_3 = %User{name: "Jane", age: 28} user_4 = %User{name: "Joe", age: 60, type: :administrator} user_5 = %User{name: "Jelly", age: 15} list_of_users = [user_1, user_2, user_3, user_4, user_5] # [ # %User{age: 25, name: "John", type: :regular}, # %User{age: 30, name: "Doris", type: :administrator}, # %User{age: 28, name: "Jane", type: :regular}, # %User{age: 60, name: "Joe", type: :administrator}, # %User{name: "Jelly", age: 15, type: :regular} # ] count = User.count_older_than(list_of_users, 28) # 2 count # 2 count = User.count_older_than(list_of_users, 25) # 3 count = User.count_older_than(list_of_users, 20) # 4
Now lets add these functions to the previous module and copy it again to iex:
def extract_administrators(list) when is_list(list) do extract_administrators(list, []) end def extract_administrators([], acc), do: acc def extract_administrators([%User{type: :administrator} = user | t], acc) do extract_administrators(t, [user | acc]) end def extract_administrators([_| t], acc), do: extract_administrators(t, acc)
And then run
admins = User.extract_administrators(list_of_users) # [ # %User{age: 60, name: "Joe", type: :administrator}, # %User{age: 30, name: "Doris", type: :administrator} # ]
So you can see that we called first the
extract_administrators with only 1 argument, a list, so it matched the first function definition. This function all it did was call the same function, but now with two arguments, the second being an empty list.
This is a fairly regular thing to do, that 2nd argument (in this case) is what is usually called the "accumulator" and it's a simple way of recursively calling functions and "accumulate" the results of each call. In this case it's used to build a new list with all the administrators we find in the original list.
So this call ends up as (given the list we were working with):
extract_administrators([%User{name: "John", age: 25, type: :regular} | t], [])
Given that the first argument is not an empty list, it can't match the second function declaration. Given that the first cell in the first argument list is not a
%User{} struct with the type
:administrator it can't match the 3rd function, so it matches the 4th function:
def extract_administrators([_| t], acc), do: extract_administrators(t, acc)
In this function, what happens is, we ignore that first value, and we're only interested in the remaining list.
[%User{age: 25, name: "John", type: :regular} | #head [%User{age: 30, name: "Doris", type: :administrator} | [%User{age: 28, name: "Jane", type: :regular} | [%User{age: 60, name: "Joe", type: :administrator} | [%User{name: "Jelly", age: 15, type: :regular} | [] ] ] ] ] ]
So in this case we ignore
_ the head, and we call again the function
extract_administrators with the remaining list and whatever is in the
acc variable, so, which ends up being this call:
extract_administrators( [%User{name: "Doris"...} | [%User{name: "Jane"...} | [%User{name: "Joe"...} | [%User{name: "Jelly"...} | []] ] ] ], [])
Now when this function is called with these new parameters it will actually match the 3rd function clause
def extract_administrators([%User{type: :administrator} = user | t], acc) do extract_administrators(t, [user | acc]) end
So here what we do is bind
user to the
element, and then call the function again with its tail, while also adding the user into the accumulator in the last argument.
You can see that it's a bit of symmetric operation when the head matches our constraints/pattern, we take the
head from one list, and we place that as the
head of the accumulator, then we pass the
tail of the list, from which we took the
head from, as the new list to the function.
The first time, acc was empty,
[]
If we declare
[user | acc] , what we're declaring is:
[%User{....} | []]
If we do it again with another user
[%User{} | [%User{} | []]]
And so on.
So now it will be called as:
extract_administrators( [%User{name: "Jane"...} | [%User{name: "Joe"...} | []] ], [%{name: "Doris"...}])
Which again will match only the 4th function clause.
So then it's called as:
extract_administrators( [%User{name: "Joe"...} | [%User{name: "Jelly"...} | [] ]], [%{name: "Doris"...} | []])
Which matches on the 3rd, so it will now call as:
extract_administrators( [%User{name: "Jelly"...}] | []], [%User{name: "Joe"...} | [%{name: "Doris"...} | []]])
Which again matches only the forth it's called again this time with an empty list as the first argument.
And now because the first argument is empty, it matches on the 2nd function clause, where we end the recursion and just return the
acc argument, leading to:
[%User{name: "Joe"...}, %{name: "Doris"...}]
These are also composable, for instance you can use:
list_of_users |> User.extract_administrators() |> User.count_older_than(28) # 2
The last part of this write up is about using function clauses with different arities and pattern-matches in anonymous functions. There are a lot of modules in the standard lib that take functions as one of their parameters, specially those that deal with collections, such as the
Enum module.
The
Enum module has one function called
reduce/3 that basically is an abstraction over what we did we these functions. It takes an enumerable, an accumulator and a function, and we can use it to reduce the elements of the enumerable into whatever accumulator we want. We could write the
extract_administrators function as:
def extract_administrators(list) when is_list(list) do Enum.reduce(list, [], fn (%User{type: :administrator} = user, acc) -> [user | acc] (_, acc) -> acc end) end
You can see two clauses on the
fn declaration.
Enum.reduce passes one element at a time from the
enumerable provided as the first argument, to the anonymous function, along with the 2nd argument as the accumulator. The first time the anonymous function is called, the accumulator is the original accumulator in the 2nd argument (an empty list), and on the following ones it's whatever the anonymous function returned.
Since we receive each element one by one (outside of their original list) we just pattern match on the element. When all elements from the enumerable have been
enumerated it returns whatever is the value of the accumulator.
There are plenty of use cases for recursive traversal of collections, in functional languages that's usually how you work on collections of items. There are also other useful functions in the
Enum module. I use
reduce and
map a lot but there are more.
There's also other tricks you can use, such as using the same name for variables, underneath this makes the pattern only succeed if all instances of the binding resolve to the same value.
So for instance let's say you wanted to take pairs of users out of that list, that shared the same type of user.
def pluck_pairs(list, type) when is_list(list) do pluck_pair(list, type, {nil, []}) end def pluck_pairs([], type, full_acc), do: full_acc def pluck_pairs([%{type: type} = user | t], type, {nil, acc}) do pluck_pairs(t, type, {user, acc}) end def pluck_pairs([%{type: type} = user | t], type, {previous, acc}) do pluck_pairs(t, type, {nil, [{previous, user} | acc]}) end def pluck_pairs([_ | t], type, acc), do: pluck_pairs(t, type, acc)
Or using the
Enum.reduce form
def pluck_pairs(list_of_users, type) do Enum.reduce(list_of_users, {nil, []}, fn (%User{type: ^type} = user, {nil, acc}) -> {user, acc} (%User{type: ^type} = user, {previous_match, acc}) -> {nil, [{previous_match, user} | acc]} (_, full_acc) -> full_acc end) end
(we need to use the
^ pin operator to pin down the value of
type, otherwise it would be re-bound during the
Enum.reduce and match anything)
Another thing to keep in mind is, if the order of the accumulation matters, that after reducing a collection the accumulator will have the elements in the inverse order, so if that's relevant you need to reverse the list, like with
Enum.reverse/1 or the erlang
:lists.reverse/1 (if it's a lit obviously) function.
And now you could use this inside a case function
case User.pluck_pairs(list_of_users, :regular) do {nil, [_|_] = acc} -> IO.puts("No unmatched user and #{length(acc)} matched pairs") Notifications.send_pair_emails(acc) {%{name: name} = no_pair, [_|_] = acc} -> IO.puts("Unmatched user #{name} and #{length(acc)} matched pairs") Notifications.send_no_pair_email(user) Notifications.send_pair_emails(acc) {%{name: name} = user, acc} -> IO.puts("Unmatched user #{name} and no pairs") Notifications.send_no_pair_email(user) acc {nil, acc} -> IO.puts("No pairs and no unmatched users") acc end
And so on. Given the natural support for concurrent and parallel processes in Erlang & Elixir, the usage of message passing and so on, pattern matching becomes even more useful, as it's fairly straightforward to describe state-machine'y behaviours using a combination of processes, receive blocks and pattern matching. Of course, most of that is already quite abstracted into higher OTP constructs such as gen_server, gen_statem and other friends. Hope this post helped you understand better pattern matching and illustrated some use cases, although there's plenty more that can't simply be covered in detail here - nonetheless try out and experiment! | https://micaelnussbaumer.com/2019/05/11/pattern-matching-in-elixir.html | CC-MAIN-2021-25 | refinedweb | 5,553 | 58.05 |
.
Terry
--------------------------
Simple output re-direction by redefining `overflow'
==========================
Suppose you have a function `write_to_window' that writes characters
to a `window' object. If you want to use the ostream function to write
to it, here is one (portable) way to do it. This depends on the
default buffering (if any).
#include <iostream.h>
/* Returns number of characters successfully written to";
}
This is Info file iostream.info, produced by Makeinfo-1.55 from the
input file ./iostream.texi.
START-INFO-DIR-ENTRY
* iostream: (iostream). The C++ input/output facility.
END-INFO-DIR-ENTRY
This file describes libio, the GNU library for C++ iostreams and C
stdio.
libio includes software developed by the University of California,
Berkeley..
Typically, sync() involves writing out any elements between the beginning and next pointers for the output buffer. It does not involve putting back any elements between the next and end pointers for the input buffer. If the function cannot succeed, it returns -1. The default behavior is to return zero.
virtual int_type underflow();
function underflow() extracts the current element c from the input stream, without advancing the current stream position.
c is the element stored in the read position.
typedef T traits_type
virtual int_type overflow(int_type c = T::eof());
If c does not compare equal to T::eof(), the function inserts the element ::to_char_type(c) into the output stream.
The example is given in my previous answer
Experts Exchange Solution brought to you by
Facing a tech roadblock? Get the help and guidance you need from experienced professionals who care. Ask your question anytime, anywhere, with no hassle.Start your 7-day free trial | https://www.experts-exchange.com/questions/10038227/Overloading-streambuf.html | CC-MAIN-2018-43 | refinedweb | 268 | 50.94 |
Hey if someone could help me shed some light on my code as too why it is getting the error String index out of range: -1 i would be for ever in your debt
the part of the question im stuck on is this
"At the beginning, your program should ask the user to specify a one-to-one mapping between the
characters in the above mentioned alphabet. For example, your program may display the 40 characters
in one line and ask the user to type the same set of characters in the second line, but in a different
order. After the mapping is established, your program should then ask the user to type in a sentence
using the characters from this alphabet, and then output the encrypted sentence"
The bit in red is the bit im really stuck on
Code:
package question3;
import java.util.Scanner;
public class Question3 {
private String alpha = "ABCDEFGHIJKLMNOPQRSTUVWXYZ1234567890 ,.!";
private String MAP = "";
private String Pass = "";
Scanner kb = new Scanner(System.in);
public void getString(){
System.out.println("These are the Characters that are able to be used");
System.out.println(alpha);
System.out.println("Please input encryption");
MAP = kb.nextLine();
MAP = MAP.toUpperCase();
}
public void checkalpha(){
int firstIndex = 0;
int lastIndex = 0;
int count = 1;
boolean repeatedChars = false;
if(MAP.length() == 40){
for (int i = 0; count < MAP.length(); i++){
char character = MAP.charAt(i);
firstIndex = MAP.indexOf(character);
lastIndex = MAP.lastIndexOf(character);
if(firstIndex != lastIndex) {
repeatedChars = true;
}
count++;
}
if(repeatedChars == true) {
System.out.println("There were repeated characters in the string");
System.out.println("Please Do not Input the characters more than Once");
}
else {
System.out.println("No Repeated Characters");
}
if (MAP.equals(alpha)){
System.out.println("Please Do not repeat the starting characters");
getString();
}
}
else {
System.out.println("Incorrect Amount of Characters");
System.out.println("");
}
}
public void printConverted(){
char PassIndex;
System.out.println("Please Input message wanted to be converted");
Pass = kb.nextLine();
for (int i = 0; i < Pass.length(); i++){
PassIndex = MAP.charAt(Pass.indexOf(alpha.charAt(i)));
System.out.println(PassIndex);
}
}
printConverted method is the one with the problem, im positive about that.
Thanks for taking a look, if you need more info about something please let me know, i am very new to java so i really want to learn what im doing wrong | https://www.javaprogrammingforums.com/whats-wrong-my-code/17674-need-some-help-my-code.html | CC-MAIN-2020-29 | refinedweb | 384 | 50.43 |
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In a Haskell-like language, could we unify map :: (a -> b) -> [a] -> [b] and mapM :: (Monad m) => (a -> b) -> [a] -> m [b] (and all the similar pairs like filter/filterM, fold/foldM, ...) by adopting the monadic definitions with the "normal" (suffix-free) names and extending the type system to allow implicit conversion into and out of the Identity monad?
map :: (a -> b) -> [a] -> [b]
mapM :: (Monad m) => (a -> b) -> [a] -> m [b]
It seems that such conversions aren't as problematic as a general subtyping system, since proving confluence isn't a problem. We'd have to do something to prevent the typechecker from looping indefinitely trying to unify a with Identity a by way of Identity (Identity (Identity (Identity (... a ...)))), but I'm not sure that this would be a problem in practice. (Perhaps by avoiding an explicit transitivity-of-subtyping deduction rule, keeping the Identity-introduction, Identity-elimination, and reflexivity rules as the only deduction rules, and proving that transitivity follows from this?)
a
Identity a
Identity (Identity (Identity (Identity (... a ...))))
Are there other difficulties that this approach might cause? What sort of trickery might be needed to convince an optimizer over a typed intermediate language that the Identity introduction and elimination functions should have no operational effect? (Furthermore, is it true that they should have no effect, or is there a useful distinction between _|_ :: Identity a and Identity _|_ :: Identity a?)
_|_ :: Identity a
Identity _|_ :: Identity a
You can do some fun tricks with type families. Then you can create true identity and composition functors.
The problem is that you have to help the type checker, because it can't tell which functor you mean when you say a -> a, it could be Identity, or the composition of Identity and Identity etc.
Perhaps playing with this code might give some insights in what problems you might encounter when constructing a language like you propose.
{-# LANGUAGE TypeFamilies, TypeOperators, UndecidableInstances #-}
import Prelude hiding (Functor(..), Monad(..))
-- Functor definition
type family F f a :: *
class Functor f where
fmap :: f -> (a -> b) -> F f a -> F f b
-- Identity functor
data Id = Id
type instance F Id a = a
instance Functor Id where
fmap Id f = f
-- Functor composition
data f :.: g = f :.: g
type instance F (f :.: g) a = F f (F g a)
instance (Functor f, Functor g) => Functor (f :.: g) where
fmap (f :.: g) = fmap f . fmap g
test1 = fmap (Id :.: Id) (+1) 2
class Functor m => Monad m where
return :: m -> a -> F m a
bind :: m -> F m a -> (a -> F m b) -> F m b
instance Monad Id where
return Id a = a
bind Id a f = f a
test2 =
bind Id 1 $ \x ->
bind Id 2 $ \y ->
return Id (x + y)
That's a great idea, Sjoerd, thanks a lot!
If you have arity-polymorphism (as well described in Arity-Generic Datatype-Generic Programming) at the level of type constructors (which is not described in the above paper!), then these unifications are less troublesome. Hopefully someone will do that particular combination soon - I could use it. | http://lambda-the-ultimate.org/node/3842 | CC-MAIN-2021-39 | refinedweb | 539 | 50.46 |
GCC Bugs The latest version of this document is always available at [1]. _________________________________________________________________ Table of Contents * [2]Reporting Bugs + [3]What we need + [4]What we DON'T want + [5]Where to post it + [6]Detailed bug reporting instructions + [7]Detailed bug reporting instructions for GNAT + [8]Detailed bug reporting instructions when using a precompiled header * [9]Frequently Reported Bugs in GCC + [10]C++ o [11]Missing features o [12]Bugs fixed in the 3.4 series + [13]Fortran * [14]Non-bugs + [15]General + [16]C + [17]C++ o [18]Common problems when upgrading the compiler _________________________________________________________________ Reporting Bugs The main purpose of a bug report is to enable us to fix the bug. The most important prerequisite for this is that the report must be complete and self-contained, which we explain in detail below. Before you report a bug, please check the [19]list of well-known bugs and, if possible in any way, try a current development snapshot. If you want to report a bug with versions of GCC before 3.1 we strongly recommend upgrading to the current release first. Before reporting that GCC compiles your code incorrectly, please compile it with gcc -Wall and see whether this shows anything wrong with your code that could be the cause instead of a bug in GCC. Summarized bug reporting instructions After this summary, you'll find detailed bug reporting instructions, that explain how to obtain some of the information requested in this summary. What we need Please include in your bug report all of the following items, the first three of which can be obtained from the output of gcc -v: * the exact version of GCC; * the system type; * the options given when GCC was configured/built; * the complete command line that triggers the bug; * the compiler output (error messages, warnings, etc.); and * the preprocessed file (*.i*) that triggers the bug, generated by adding -save-temps to the complete compilation command, or, in the case of a bug report for the GNAT front end, a complete set of source files (see below). the compiler? :-) * An error that occurs only some of the times a certain file is compiled, such that retrying a sufficient number of times results in a successful compilation; this is a symptom of a hardware problem, not of a compiler bug (sorry) * * Bugs in releases or snapshots of GCC not issued by the GNU Project. 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G++ emits two copies of constructors and destructors. In general there are three types of constructors (and destructors). 1. The complete object constructor/destructor. 2. The base object constructor/destructor. 3. The allocating constructor/deallocating destructor. The first two are different, when virtual base classes are involved.). Classes in exception specifiers must be complete types. [15.4]/1 tells you that you cannot have an incomplete type, or pointer to incomplete (other than cv void *) in an exception specification. Exceptions don't work in multithreaded applications. You need to rebuild g++ and libstdc++ with --enable-threads. Remember, C++ exceptions are not like hardware interrupts. You cannot throw an exception in one thread and catch it in another. You cannot throw an exception from a signal handler and catch it in the main thread. Templates, scoping, and digraphs.. Copy constructor access check while initializing a reference.. Common problems when upgrading the compiler ABI changes [35. Standard conformance With each release, we try to make G++ conform closer to the ISO C++ standard (available at [36]). We have also implemented some of the core and library defect reports (available at [37] & [38]). New in GCC 3.0 *: * Say std::cout at the call. This is the most explicit way of saying what you mean. * Say using std::cout; somewhere before the call. You will need to do this for each function or type you wish to use from the standard library. * Say using namespace std; somewhere before the call. This is the quick-but-dirty fix. This brings the whole of the std:: namespace into scope. Never do this in a header file, as every user of your header file will be affected by this decision. New in GCC 3.4.0 The new parser brings a lot of improvements, especially concerning name-lookup. * The "implicit typename" extension got removed (it was already deprecated since GCC 3.1), so that the following code is now rejected, see [14.6]: template <typename> struct A { typedef int X; }; template <typename T> struct B { A<T>::X x; // error typename A<T>::X y; // OK }; B<void> b; * For similar reasons, the following code now requires the template keyword, see [14.2]: template <typename> struct A { template <int> struct X {}; }; template <typename T> struct B { typename A<T>::X<0> x; // error typename A<T>::template X<0> y; // OK }; B<void> b; * We now have two-stage name-lookup, so that the following code is rejected, see [14.6]/9: template <typename T> int foo() { return i; // error } * This also affects members of base classes, see [14.6.2]: [39]Common Misunderstandings with GNU C++. References 1. 2. 3. 4. 5. 6. 7. 8. 9. 10. 11. 12. 13. 14. 15. 16. 17. 18. 19. 20. 21. mailto:gcc-bugs@gcc.gnu.org 22. 23. 24. 25. 26. 27. 28. 29. 30. 31. 32. 33. 34. 35. 36. 37. 38. 39. | http://opensource.apple.com//source/libstdcxx/libstdcxx-11/libstdcxx/BUGS | CC-MAIN-2016-44 | refinedweb | 1,462 | 63.09 |
filevault 0.1.0
filevault: manage a tree of directories and files
filevault
filevault is a class for managing a tree of files on the filesystem.
A Vault will:
- Create a directory tree of custom depth
- Spreads files out which keeps CLI snappy when traversing the tree
- Scale to hundreds of thousands of files
- Obfuscate directory paths and filenames
how to install
Setuptools:
easy_install filevault
Pip:
pip install filevault
how to use
How to create a default Vault object:
from filevault import Vault v = Vault()
You may (and should) customize the Vault instance. Here are the arguments:
- vaultpath
- Where should the tree be created? Defaults to ‘vault’ in pwd.
- depth
- How deep should the tree span? Defaults to 3 directories deep.
- salt
- Add a custom salt for a unique and more secure tree, defaults to ‘changeme’
Another custom example:
from filevault import Vault v = Vault( vaultpath="/tmp/for-test", depth=2, salt="sugar" )
Now that we have a Vault object named v, we may review its two methods:
- create_filename( seed, ext=” )
- Create a valid vault filename seeded with input. Optional extension.
- create_random_filename( ext=” )
- Create a valid vault filename seeded with random input. Optional extension.
Here is a full example:
# import Vault class from filevault import Vault # create vault object named v, with custom path, depth, and salt v = Vault( vaultpath="/tmp/for-test", depth=2, salt="sugar" ) # print a valid vault filename with extension print v.create_filename( "my-first-file", ".png" ) # result: # 3/9/3993817d4f9b3867c6db29b23c9d2ff9bb8a87d89426002adbb6ed34289d9e32.png print v.create_filename( "my-first-file", ".png" ) # Same result: # 3/9/3993817d4f9b3867c6db29b23c9d2ff9bb8a87d89426002adbb6ed34289d9e32.png # print a random valid vault filename without extension v.create_random_filename() # result: # 6/1/6169d6ee0ac0bc63ab667fb94d9cc747df0c03596ac43e24a51b3517d74bdc42
how can I thank you?
Check out my webpage to screenshot service and give me some feedback, tips, or advice. Every little bit of help counts.
- Author: Russell Ballestrini
- Keywords: file vault filevault tree directories obfuscate
- License: Public Domain
- Platform: All
- Package Index Owner: russellballestrini
- DOAP record: filevault-0.1.0.xml | https://pypi.python.org/pypi/filevault/0.1.0 | CC-MAIN-2017-22 | refinedweb | 325 | 53.92 |
I am trying to duplicate the example rhino-inside Convert ConsoleApp written in C# into Python, but really having no luck.
Sorry if there’s an obvious solution, but the documentation does not seem to match up to reality when trying to use Python.
The C# app is creating a ‘RhinoCore’ but I have not been able to figure out how to get it within Python, nor how to ‘Create’ a RhinoDoc, since it seems to want some unspecified argument(s).
The example code is:
import rhinoinside rhinoinside.load() import System import Rhino fn = "c:\\drawing.stl" cmd = "_import \"{0}\"".format(fn) doc = Rhino.RhinoDoc.Create() opts = Rhino.FileIO.FileObjReadOptions(Rhino.FileIO.FileReadOptions()) obj = Rhino.FileIO.FileObj.Read(cmd, doc, opts)
It fails on the Rhino.RhinoDoc.Create() wanting some unspecified argument, but the documentation does not show any. Not sure where to go next. Anybody have any ideas?
PS: the error message is:
PS C:\Users\krivacic\Rhino_Tests\Test1> python .\example.py Traceback (most recent call last): File ".\example.py", line 10, in <module> doc = Rhino.RhinoDoc.Create() TypeError: No method matches given arguments | https://discourse.mcneel.com/t/rhino-inside-and-python/82457 | CC-MAIN-2019-39 | refinedweb | 185 | 53.68 |
[
]
Sergi Vladykin commented on IGNITE-2906:
----------------------------------------
For the issue you've described you can use @QuerySqlField(name = "unique_id") annotation.
Probably the issue description is wrong.
Also you should provide respective unit tests in your patch.
> Embedded / child element types indexing/queryfields (non-flat)
> --------------------------------------------------------------
>
> Key: IGNITE-2906
> URL:
> Project: Ignite
> Issue Type: Improvement
> Components: cache, data structures, general, SQL
> Reporter: Krome Plasma
> Attachments: indexing.patch
>
>
> I've had discussion about this on Apache Ignite Users.
>
> The problem occurs when you want to index a non-primitive type that have same names of
variables as the enclosing type, better described on forum above. As a short example:
> Let's say we want to index:
> public class Person
> {
> @QuerySqlField
> long id;
> @QuerySqlField
> PersonData personData;
> }
> public class PersonData
> {
> @QuerySqlField
> long id;
> }
> This will not work as it will detect indexes/query fields with same names for index Person.id
and PersonData.id because the names are flattened to simply 'id'.
> I am attaching a simple patch that resolves this issue. We've been running this for (3
months now) and found no problems. However we are using annotations and not XML. I am not
sure the patch completely solves the problem for XML based configuration.
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