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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import matplotlib.pyplot as plt import pandas as pd from matplotlib import rcParams import matplotlib.cm as cm import matplotlib as mpl #colorbrewer2 Dark2 qualitative color table)] rcParams['figure.figsize'] = (10, 6) rcParams['figure.dpi'] = 150 rcParams['axes.color_cycle'] = dark2_colors rcParams['lines.linewidth'] = 2 rcParams['axes.facecolor'] = 'white' rcParams['font.size'] = 14 rcParams['patch.edgecolor'] = 'white' rcParams['patch.facecolor'] = dark2_colors[0] rcParams['font.family'] = 'StixGeneral' def remove_border(axes=None, top=False, right=False, left=True, bottom=True): """ Minimize chartjunk by stripping out unnecesasry() pd.set_option('display.width', 500) pd.set_option('display.max_columns', 100)
#this mapping between states and abbreviations will come in handy later states_abbrev = { ', 'NA': ' }.
Your Answer Here?
Your Answer Here
Now, we run the simulation with this model, and plot it.
model = simple_gallup_model(gallup_2012) model = model.join(electoral_votes) prediction = simulate_election(model, 10000) plot_simulation(prediction) plt.show() make_map(model.Obama, "P(Obama): Simple Model")
1.7 Attempt to validate the predictive model using the above simulation histogram. Does the evidence contradict the predictive model?
Your answer here.
Your answers here
1.11 Simulate elections assuming a bias of 1% and 5%, and plot histograms for each one.
#your code
A quick look at the graph will show you a number of states where Gallup showed a Democratic advantage, but where the elections were lost by the democrats. Use Pandas to list these states.
#your code here
We compute the average difference between the Democrat advantages in the election and Gallup poll
print (prediction_08.Dem_Adv - prediction_08.Dem_Win).mean()
your answer here
1.13 Calibrate your forecast of the 2012 election using the estimated bias from 2008. Validate the resulting model against the real 2012 outcome. Did the calibration help or hurt your prediction?
#your code here
1.14 Finally, given that we know the actual outcome of the 2012 race, and what you saw from the 2008 race would you trust the results of the an election forecast based on the 2012 Gallup party affiliation poll?
Your answer here
your answer here
your answer here
2.5 As before, plot a histogram and map of the simulation results, and interpret the results in terms of accuracy and precision.
#code to make the histogram #your code here
#code to make the map #your code here
your answer here
your answer here
3.3 Run 10,000 simulations with this model, and plot the results. Describe the results in a paragraph -- compare the methodology and the simulation outcome to the Gallup poll. Also plot the usual map of the probabilities
#your code here
Your summary here
#your code here
Not all polls are equally valuable. A poll with a larger margin of error should not influence a forecast as heavily. Likewise, a poll further in the past is a less valuable indicator of current (or future) public opinion. For this reason, polls are often weighted when building forecasts.
A weighted estimate of Obama's advantage in a given state is given by
$$ \mu = \frac{\sum w_i \times \mu_i}{\sum w_i} $$
where $\mu_i$ are individual polling measurements or a state, and $w_i$ are the weights assigned to each poll. The uncertainty on the weighted mean, assuming each measurement is independent, is given by
The estimate of the variance of $\mu$, when $\mu_i$ are unbiased estimators of $\mu$, is
$$\textrm{Var}(\mu) = \frac{1}{(\sum_i w_i)^2} \sum_{i=1}^n w_i^2 \textrm{Var}(\mu_i).$$
We need to find an estimator of the variance of $\mu_i$, $Var(\mu_i)$. In the case of states that have a lot of polls, we expect the bias in $\mu$ to be negligible, and then the above formula for the variance of $\mu$ holds. However, lets take a look at the case of Kansas.
multipoll[multipoll.State=="Kansas"]
There are only two polls in the last year! And, the results in the two polls are far, very far from the mean.
Now, Kansas is a safely Republican state, so this dosent really matter, but if it were a swing state, we'd be in a pickle. We'd have no unbiased estimator of the variance in Kansas. So, to be conservative, and play it safe, we follow the same tack we did with the unweighted averaging of polls, and simply assume that the variance in a state is the square of the standard deviation of
obama_spread.
This will overestimate the errors for a lot of states, but unless we do a detailed state-by-state analysis, its better to be conservative. Thus, we use:
$\textrm{Var}(\mu)$ =
obama_spread.std()$^2$ .
The weights $w_i$ should combine the uncertainties from the margin of error and the age of the forecast. One such combination is:
$$ w_i = \frac1{MoE^2} \times \lambda_{\rm age} $$
where
$$ \lambda_{\rm age} = 0.5^{\frac{{\rm age}}{30 ~{\rm days}}} $$
This model makes a few ad-hoc assumptions:
3.4 Nevertheless, it's worth exploring how these assumptions affect the forecast model. Implement the model in the function
weighted_state_average
""" Function -------- weighted
3.5 Put this all together -- compute a new estimate of
poll_mean and
poll_std for each state, apply the
default_missing function to handle missing rows, build a forecast with
aggregated_poll_model, run 10,000 simulations, and plot the results, both as a histogram and as a map.
#your code here
#your map code here make_map(model.Obama, "P(Obama): Weighted Polls")
3.6 Discuss your results in terms of bias, accuracy and precision, as before
your answer here
For fun, but not to hand in, play around with turning off the time decay weight and the sample error weight individually.
The models we have explored in this homework have been fairly ad-hoc. Still, we have seen predicting by simulation, prediction using heterogeneous side-features, and finally by weighting polls that are made in the election season. The pros pretty much start from poll-averaging, adding in demographics and economic information, and moving onto trend-estimation as the election gets closer. They also employ models of likely voters vs registered voters, and how independents might break. At this point, you are prepared to go and read more about these techniques, so let us leave you with some links to read:
Skipper Seabold's reconstruction of parts of Nate Silver's model: . We've drawn direct inspiration from his work , and indeed have used some of the data he provides in his repository
The simulation techniques are partially drawn from Sam Wang's work at . Be sure to check out the FAQ, Methods section, and matlab code on his site.
Nate Silver, who we are still desperately seeking, has written a lot about his techniques: . Start there and look around
Drew Linzer uses bayesian techniques, check out his work at:
How to submit
To submit your homework, create a folder named lastname_firstinitial_hw2 and place this notebook file in the folder. Also put the data folder in this folder. Make sure everything still works! Select Kernel->Restart Kernel to restart Python, Cell->Run All to run all cells. You shouldn't hit any errors. Compress the folder (please use .zip compression) and submit to the CS109 dropbox in the appropriate folder. If we cannot access your work because these directions are not followed correctly, we will not grade your work.
css tweaks in this cell | http://nbviewer.jupyter.org/github/cs109/content/blob/master/HW2.ipynb | CC-MAIN-2018-51 | refinedweb | 1,207 | 57.37 |
In this article, we will learn how to find all matches to the regular expression in Python. The RE module’s
re.findall() method scans the regex pattern through the entire target string and returns all the matches that were found in the form of a Python list.
Table of contents
- How to use re.findall()
- Example to find all matches to a regex pattern
- finditer method
- Regex find all word starting with specific letters
- Regex to find all word that starts and ends with a specific letter
- Regex to find all words containing a certain letter
- Regex findall repeated characters
How to use re.findall()
Before moving further, let’s see the syntax of the
re.findall() method.
Syntax
re.findall(pattern, string, flags=0)
pattern: regular expression pattern we want to find in the string or text
string: It is the variable pointing to the target string (In which we want to look for occurrences of the pattern).
Flags: It refers to optional flags. by default, no flags are applied. For example, the
re.Iflag is used for performing case-insensitive findings.
The regular expression pattern and target string are the mandatory arguments, and flags are optional.
Return Value
The
re.findall() scans the target string from left to right as per the regular expression pattern and returns all matches in the order they were found.
It returns
None if it fails to locate the occurrences of the pattern or such a pattern doesn’t exist in a target string.
Example to find all matches to a regex pattern
In this example, we will find all numbers present inside the target string. To achieve this, let’s write a regex pattern.
Pattern:
\d+
What does this pattern mean?
- The
\dis a special regex sequence that matches any digit from 0 to 9 in a target string.
- The
+metacharacter indicates number can contain at minimum one or maximum any number of digits.
In simple words, it means to match any number inside the following target string.
target_string = "Emma is a basketball player who was born on June 17, 1993. She played 112 matches with scoring average 26.12 points per game. Her weight is 51 kg."
As we can see in the above string ’17’, ‘1993’, ‘112’, ’26’, ’12’, ’51’ number are present, so we should get all those numbers in the output.
Example
import re target_string = "Emma is a basketball player who was born on June 17, 1993. She played 112 matches with scoring average 26.12 points per game. Her weight is 51 kg." result = re.findall(r"\d+", target_string) # print all matches print("Found following matches") print(result) # Output ['17', '1993', '112', '26', '12', '51']
Note:
First of all, I used a raw string to specify the regular expression pattern i.e
r"\d+". As you may already know, the backslash has a special meaning in some cases because it may indicate an escape character or escape sequence to avoid that we must use raw string.
finditer method
The
re.finditer() works exactly the same as the
re.findall() method except it returns an iterator yielding match objects matching the regex pattern in a string instead of a list. It scans the string from left-to-right, and matches are returned in the iterator form. Later, we can use this iterator object to extract all matches.
In simple words,
finditer() returns an iterator over MatchObject objects.
But why use
finditer()?
In some scenarios, the number of matches is high, and you could risk filling up your memory by loading them all using
findall(). Instead of that using the
finditer(), you can get all possible matches in the form of an iterator object, which will improve performance.
It means,
finditer() returns a callable object which will load results in memory when called. Please refer to this Stackoverflow answer to get to know the performance benefits of iterators.
finditer example
Now, Let’s see the example to find all two consecutive digits inside the target string.
import re target_string = "Emma is a basketball player who was born on June 17, 1993. She played 112 matches with a scoring average of 26.12 points per game. Her weight is 51 kg." # finditer() with regex pattern and target string # \d{2} to match two consecutive digits result = re.finditer(r"\d{2}", target_string) # print all match object for match_obj in result: # print each re.Match object print(match_obj) # extract each matching number print(match_obj.group())
Output:
re.Match object; span=(49, 51), match='17' 17 re.Match object; span=(53, 55), match='19' 19 re.Match object; span=(55, 57), match='93' 93 re.Match object; span=(70, 72), match='11' 11 re.Match object; span=(103, 105), match='26' 26 re.Match object; span=(106, 108), match='12' 12 re.Match object; span=(140, 142), match='51' 51
More use
- Use finditer to find the indexes of all regex matches
- Regex findall special symbols from a string
Regex find all word starting with specific letters
In this example, we will see solve following 2 scenarios
- find all words that start with a specific letter/character
- find all words that start with a specific substring
Now, let’s assume you have the following string:
target_string = "Jessa is a Python developer. She also gives Python programming training"
Now let’s find all word that starts with letter p. Also, find all words that start with substring ‘py‘
Pattern:
\b[p]\w+\b
- The
\bis a word boundary, then p in square bracket
[]means the word must start with the letter ‘p‘.
\w+means one or more alphanumerical characters after a letter ‘p’
- In the end, we used
\bto indicate word boundary i.e. end of the word.
Example
import re target_string = "Jessa is a Python developer. She also gives Python programming training" # all word starts with letter 'p' print(re.findall(r'\b[p]\w+\b', target_string, re.I)) # output ['Python', 'Python', 'programming'] # all word starts with substring 'Py' print(re.findall(r'\bpy\w+\b', target_string, re.I)) # output ['Python', 'Python']
Regex to find all word that starts and ends with a specific letter
In this example, we will see solve following 2 scenarios
- find all words that start and ends with a specific letter
- find all words that start and ends with a specific substring
Example
import re target_string = "Jessa is a Python developer. She also gives Python programming training" # all word starts with letter 'p' and ends with letter 'g' print(re.findall(r'\b[p]\w+[g]\b', target_string, re.I)) # output 'programming' # all word starts with letter 'p' or 't' and ends with letter 'g' print(re.findall(r'\b[pt]\w+[g]\b', target_string, re.I)) # output ['programming', 'training'] target_string = "Jessa loves mango and orange" # all word starts with substring 'ma' and ends with substring 'go' print(re.findall(r'\bma\w+go\b', target_string, re.I)) # output 'mango' target_string = "Kelly loves banana and apple" # all word starts or ends with letter 'a' print(re.findall(r'\b[a]\w+\b|\w+[a]\b', target_string, re.I)) # output ['banana', 'and', 'apple']
Regex to find all words containing a certain letter
In this example, we will see how to find words that contain the letter ‘i’.
import re target_string = "Jessa is a knows testing and machine learning" # find all word that contain letter 'i' print(re.findall(r'\b\w*[i]\w*\b', target_string, re.I)) # found ['is', 'testing', 'machine', 'learning'] # find all word which contain substring 'ing' print(re.findall(r'\b\w*ing\w*\b', target_string, re.I)) # found ['testing', 'learning']
Regex findall repeated characters
For example, you have a string:
""Jessa Erriika""
As the result you want to have the following matches:
(J, e, ss, a, E, rr, ii, k, a)
Example
import re target_string = "Jessa Erriika" # This '\w' matches any single character # and then its repetitions (\1*) if any. matcher = re.compile(r"(\w)\1*") for match in matcher.finditer(target_string): print(match.group(), end=", ") # output J, e, ss, a, E, rr, ii, k, a, | https://pynative.com/python-regex-findall-finditer/ | CC-MAIN-2021-10 | refinedweb | 1,343 | 65.62 |
ESLint works with a set of rules you define. Our basic configuration contains just one such rule: strings should be written inside single quotes rather than double quotes. You can add more if you want, but it's more common to find an existing set of rules that come close to what you want, then customise from there.
Arguably the most common linting rules around are by Airbnb, which bills itself as "a mostly reasonable approach to JavaScript." And it's true: their linting rules are popular because they are simple, sensible, and beautifully consistent.
We're going to install their Airbnb rules for ESLint and see what it makes of our source code. Run this command in your terminal window:
npm install --save-dev eslint-config-airbnb
We now just need to tell ESLint that our own rules extend their Airbnb rules. This uses the Airbnb rules as a foundation, adding our own overrides as needed. Modify your .eslintrc file to this:
.eslintrc
{ "parser": "babel-eslint", "env": { "browser": true, "node": true }, "extends": "airbnb", "rules": { "indent": [2, "tab"] } }
There's still only one rule in there, but I've modified it to something deeply contentious because we're almost at the end now so I feel it's safe to take some risks. This new rule means "make sure I use tabs for indenting rather than spaces," and if that doesn't give you enough motivation to search for ESLint configuration options, I don't know what will! (Note: if you either don't want tabs or don't want to figure out how to set something else in the linter options, just delete the rule.)
Anyway, save your new configuration file and run
npm run lint in your terminal window. This time you'll see lots of errors fill your screen, all telling you what the problem was as well as a filename and line number. Note that these errors are all stylistic rather than actual bugs, but like I said it's important to fix these issues too if you want clear, clean, maintainable code.
Let's tackle the easy ones first, starting with "Newline required at end of file but not found". You might see this one a few times, and it's trivial to fix: just add a blank line to the end of every file where you see this warning. Another easy one is "Missing trailing comma," which just means that code like this:
this.state = { events: [] };
…needs to be rewritten to this:
this.state = { events: [], };
The extra comma doesn't add much, but it does reduce the chance of you adding more properties without first adding a comma. Warning: don't do this in JSON files such as package.json, because many parsers will be deeply unhappy.
There are two more easy ones to fix if we choose. First, "Unexpected console statement" just means ESLint doesn't want us to use
console.log() in our code, but this is only a warning not an error so I'm happy to ignore this in my own code – it's down to you if you want to remove them in yours. Second, "'Link' is defined but never used" in User.js. To fix this problem, change this line:
src/pages/User.js
import { IndexLink, Link } from 'react-router';
…to this:
src/pages/User.js
import { IndexLink } from 'react-router';
Unless your code is very different from mine, that should fix all the easy linter errors. Now it's time for the harder! | http://www.hackingwithreact.com/read/1/40/linting-react-using-airbnbs-eslint-rules | CC-MAIN-2019-04 | refinedweb | 586 | 78.28 |
Distributing class instance indices
On 01/05/2015 at 05:41, xxxxxxxx wrote:
Hi all,
Just for the sake of clarity another thread...
Is it possible to set a range of hexadecimal indices for class instances and distribute respectively overwrite the content myself?
Thanks for any help in advance
Martin
On 04/05/2015 at 01:39, xxxxxxxx wrote:
Hi Martin,
I'm afraid you need to give us some more info. Even after discussing with Sebastian, we have no idea, what you are after. Sorry.
On 04/05/2015 at 03:05, xxxxxxxx wrote:
Hi Andreas,
I´m not quite sure if I understand the class-index concept the right way.
Cinema4d builds an index for every object instance and I thought about accessing these instances through this index.
For example, when my plugin becomes active, it occupies a set of indices in c4d which I can assign to several types of subclasses. For example, every conifer type subclass instance has an index with a 4000 in it, every broadleaf with 400 and so on.
Later on I want to access the types by calling indices.
Best wishes
Martin
On 04/05/2015 at 17:23, xxxxxxxx wrote:
Hi Martin,
Andreas is away at FMX this week, so I'm filling in on this topic. I have a rough idea what you're asking about, but could you please post some example code that accesses the data you're discussing? It would probably give me a clue as to where to search in the code to give you a solid answer.
Thanks,
Joey Gaspe
SDK Support Engineer
On 05/05/2015 at 03:11, xxxxxxxx wrote:
Hi Joey,
thanks for your help!
To keep it simple, I wrote a small pseudo code.
At the validation point inside the Plant class c4d assigns a hexadecimal index to the instances.
I want to assign those indices by myself according to their position in space, age and other attributes,
in order to store them in a hyperfile.
Later on I want to access the hyperfile and modify the Instances.
For example bend them if the wind is blowing from the west, that the "west plants" are more influenced.
Hope you can catch the drift?
Please let me know if you need more information.
Thanks in advance
Martin
#______________________________________
Pseudocode plant distribution:
Plantclass:
def init: with plant specific attributes(e.g. index, is valid, place to be, next possible places[npp])
def TestSpace() :
return if plant is placed or not valid
testing the next possible places to live in
for next possible places:
if place is valid:
place actual plant
built Plantclass instances for next possible places
instance.TestSpace() #recursively call
Forestclass:
placedplants = []
def init: with calling TestSpace from a certain start point plant.
with calling Collect(start point plant)
def Collect(self,spp) :
if plant is placed:
Forestclass.placedplants.append(plant)
for n in spp.npp:
if n != None:
Forestclass.Collect(n)#recursively call
On 11/05/2015 at 09:06, xxxxxxxx wrote:
Hi Martin,
this makes me feel bad... I (or actually we, as we discussed this in our team) still don't get it. For example I'm not sure, where Plantclass gets derived from and who's calling init... I'm not aware of a situation, where C4D assigns any indeces. But this may be due to a lack of knowledge on my side.
If it makes it easier in any way, you may also contact e via PM in German. We can then wrap this up for the forum community in the end.
On 11/05/2015 at 12:48, xxxxxxxx wrote:
Hi Andreas,
thanks for caring.
I´ll send a PM after my holidays.
Best wishes
Martin | https://plugincafe.maxon.net/topic/8686/11374_distributing-class-instance-indices/3 | CC-MAIN-2019-43 | refinedweb | 620 | 72.26 |
DESCRIPTION
This page describes the syntax of regular expressions in Perl.
If you haven't used regular expressions before, a quick-start
introduction is available in perlrequick, and a longer tutorial
introduction is available in perlretut.
For reference on how regular expressions are used in matching
operations, plus various examples of the same, see discussions of
"m//", "s///", "qr//" and "??" in "Regexp Quote-Like Operators" in
perlop.
Modifiers.
Used together, as /ms, they let the "." match any character
whatsoever, while still allowing "^" and "$" to match,
respectively, just after and just before newlines within the
string.
i Do case-insensitive pattern matching.
If "use locale" is in effect, the case map is taken from the
current locale. See perllocale.
x Extend your pattern's legibility by permitting whitespace and.
These are usually written as "the "/x" modifier", even though the
delimiter in question might not really be a slash. Any of these
modifiers may also be embedded within the regular expression itself
Perl's regular expressions more readable. Note that you have to be
careful not to include the pattern delimiter in the comment--perl has
no way of knowing you did not intend to close the pattern early. See
the C-comment deletion code in perlop. Also note that anything inside
a "\Q...\E" stays unaffected by "/x".
Regular Expressions
Metacharacters
The patterns used in Perl pattern matching evolved from those.)
To simplify multi-line substitutions, the "." character never matches a
newline unless you use the "/s" modifier, which in effect tells Perl to
pretend the string is a single line--even if it isn't.
Quantifiers. In particular, the lower bound is not optional.)
a "?". Note that the meanings don't change, just the "greediness":
*? Match 0 or more times, not greedily
+? Match 1 or more times, not greedily
?? Match 0 or 1 time, not greedily
{n}? Match exactly n times, not greedily
{n,}? Match at least n times, not greedily
{n,m}? Match at least n but not more than m times, not greedily
By default, "(?>...)" for more details;
possessive quantifiers are just syntactic sugar for that construct. For
instance the above example could also be written as follows:
/"(?>(?:(?>[^"\\]+)|\\.)*)"/
Escape sequences
Because patterns are processed as double quoted strings, the following
also work:
\t tab (HT, TAB)
\n newline (LF, NL)
\r return (CR)
\f form feed (FF)
\a alarm (bell) (BEL)
\e escape (think troff) (ESC)
\033 octal char (example: ESC)
\x1B hex char (example: ESC)
\x{263a} long hex char (example: Unicode SMILEY)
You cannot include a literal "$" or "@" within a "\Q" sequence. An
unescaped "$" or "@" interpolates the corresponding variable, while
escaping will cause the literal string "\$" to be matched. You'll need
to write something like "m/\Quser\E\@\Qhost/".
Character Classes and other Special Escapes (?>\PM\pM*)
\C Match a single C char (octet) even under Unicode.
NOTE: breaks up characters into their UTF-8 bytes,
so you may end up with malformed pieces of UTF-8.
Unsupported in lookbehind.
\1 Backreference to a specific group.
'1' may actually be any positive integer.
\g1 Backreference to a specific or previous group,
\g{-1} number may be negative indicating a previous buffer and may
optionally be wrapped in curly brackets for safer parsing.
\g{name} Named backreference
\k<name> Named backreference
\K Keep the stuff left of the \K, don't include it in $&
\v Vertical whitespace
\V Not vertical whitespace
\h Horizontal whitespace
\H Not horizontal whitespace
\R Linebreak
A "\w" matches a single alphanumeric character (an alphabetic
character, or a decimal digit) or "_",, but they aren't usable as either end of
a range. If any of them precedes or follows a "-", the "-" is
understood literally. If Unicode is in effect, "\s" matches also
"\x{85}", "\x{2028}", and "\x{2029}". See perlunicode for more details
about "\pP", "\PP", "\X" and the possibility of defining your own "\p"
and "\P" properties, and perluniintro about Unicode in general.
"\R" will atomically match a linebreak, including the network line-
ending "\x0D\x0A". Specifically, is exactly equivalent to
# this is correct:
$string =~ /[[:alpha:]]/;
# this is not, and will generate a warning:
$string =~ /[:alpha:]/;" since the "[[:space:]]" includes
also the (very rare) "vertical tabulator", "\ck" or chr(11) in
ASCII.
[3] A Perl extension, see above.
For example use "[:upper:]" to match all the uppercase characters.
Note that the "[]" are part of the "[::]" construct, not part of the
whole character class. For example:
[01[:alpha:]%]
matches zero, one, any alphabetic character, and the percent sign.
The following equivalences to Unicode \p{} constructs and equivalent
backslash character classes (if available), will hold:
[[:...:]] \p{...} backslash
alpha IsAlpha
alnum IsAlnum
ascii IsASCII
blank
cntrl IsCntrl
digit IsDigit \d
graph IsGraph
lower IsLower
print IsPrint (but see [2] below)
[1] If the "utf8" pragma is not used but the "locale" pragma is, the
classes correlate with the usual isalpha(3) interface (except for
"word" and "blank").
But if the "locale" or "encoding" pragmas are not used and the
string is not "utf8", then "[[:xxxxx:]]" (and "\w", etc.) will not
match characters 0x80-0xff; whereas "\p{IsXxxxx}" will force the
string to "utf8" and can match these characters (as Unicode).
[2] "\p{IsPrint}" matches characters 0x09-0x0d but "[[:print:]]" does
not.
[3] "[[:punct::]]" matches the following but "\p{IsPunct}" does not,
because they are classed as symbols (not punctuation) in Unicode.
"$" Currency symbol
"+" "<" "=" ">" "|" "~"
Mathematical symbols
"^" "`"
Modifier symbols (accents)
The other named classes are:
cntrl
Any control character. Usually characters that don't produce
output as such but instead control").
graph
Any alphanumeric or punctuation (special) character.
print}
. Note that the rule for zero-length matches is modified
somewhat, in that contents to the left of "\G" is not counted when
determining the length of the match. Thus the following will not match
forever:
$str = 'ABC';
pos($str) = 1;
while (/.\G/g) {
print $&;
} the warning below about \1 vs $1
for details.) Referring back to another part of the match is called a
backreference. print "'$1' is the first doubled character\n";
if (/Time: (..):(..):(..)/) { # parse out values
$hours = $1;
$minutes = $2;
$seconds = $3;
}
Several special variables also refer back to portions of the previous
perlsyn.)
NOTE: Failed matches in Perl do not reset the match variables, which
makes it easier to write code that tests for a series of more specific
cases and remembers the best match...
Extended Patterns
Perl also defines a consistent extension syntax for features not found
in standard tools like awk and lex. The syntax is a pair of
"question" exactly what is going on. That's psychology...
"(?#text)"
A comment. The text is ignored. If the "/x" modifier
enables whitespace formatting, a simple "#" will suffice.
Note that Perl closes the comment as soon as it sees a ")",
so there is no way to put a literal ")" in the comment.
"(?pimsx-imsx)"
One or more embedded pattern-match modifiers, to be turned on
(or turned off, if preceded by "-") for the remainder of the
pattern or the remainder of the enclosing pattern group (if
any). This is particularly useful for dynamic+ \1
will match "blah" in any case, some spaces, and an exact
(including the case!) repetition of the previous word,
assuming the "/x" modifier, and no "/i" modifier outside this
group.
Note that the "p" modifier is special in that it can only be
enabled, not disabled, and that its presence anywhere in a
pattern has a global effect. Thus "(?-p)" and "(?-p:...)" are
meaningless and will warn when executed under "use warnings".
"(?:pattern)"
"(?imsx-imsx:pattern)"
This is for clustering, not capturing; it groups
subexpressions like "()", but doesn't make backreferences as
"()" does. So
@fields = split(/\b(?:a|b|c)\b/)
is like
@fields = split(/\b(a|b|c)\b/)
"(?|pattern)" the captured content will be stored.
# before ---------------branch-reset----------- after
/ ( a ) (?| x ( y ) z | (p (q) r) | (t) u (v) ) ( z ) /x
# 1 2 2 3 2 3 4
Note: as of Perl 5.10.0, branch resets interfere with the
contents of the "%+" hash, that holds named captures.
Consider using "%-" instead..
"(?
tab, without including the tab in $&. Works only for
fixed-width look-behind.
There is a special form of this construct, called "\K",
which causes the regex engine to "keep" everything it had
matched prior to the "\K" and not include it in $&. This
effectively provides variable length look-behind. The use
of "\K" inside of another look-around" for are numbered sequentially regardless of
being named or not. Thus in the pattern
/(x)(?<foo>y)(z)/
$+{foo} will be the same as $2, and $3 will contain 'z'
instead of the opposite which is what a .NET regex hacker
might expect.
">".
"(?{ code })"
WARNING: This extended regular expression feature is
considered experimental, and may be changed without notice.
Code executed that has side effects may not perform
identically from version to version due to the effect of
future optimisations in the regex engine.
This zero-width assertion evaluates any embedded Perl code.
It always succeeds, and its "code" is not interpolated.
Currently, the rules to determine where the "code" ends are
somewhat convoluted.
This feature can be used together with the special variable
$^N to
within this string.
The "code" is properly scoped in the following sense: If the
assertion is backtracked (compare "Backtracking"), all
changes introduced after "local"ization are undone, so that
$_ = 'a' x 8;
m<
(?{ $cnt = 0 }) # Initialize $cnt.
(
a
(?{
local $cnt = $cnt + 1; # Update $cnt, backtracking-safe.
})
)*
the special variable $^R. This happens immediately, so $^R
can be used from other "(?{ code })" assertions inside the
same regular expression.
The assignment to $^R above is properly localized, so the old
value of $^R is restored if the assertion is backtracked;
compare "Backtracking".
Due to an unfortunate implementation issue, the Perl code
contained in these blocks is treated as a compile time
closure that can have seemingly bizarre consequences when
used with lexically scoped variables inside of subroutines or
loops. There are various workarounds for this, including
simply using global variables instead. If you are using this
construct and strange results occur then check for the use of
lexically scoped variables. due compartment. See
perlsec for details about both these mechanisms.
Because Perl's regex engine is currently not re-entrant,
interpolated code may not invoke the regex engine either
directly with "m//" or "s///"), or indirectly with functions
such as "split".
"(??{" is
evaluated at run time, at the moment this subexpression may
The "code" is not interpolated. As before, the rules to
determine where the "code" ends are
included.
The following pattern matches a function foo() which may
contain balanced parentheses as the argument.
$re = qr{ ( # paren group 1 (full function)
foo
buffer. "(condition)" should be either an
(1) (2) ...
Checks if the numbered capturing..
(?<NAME_PAT>....)
(?<ADRESS_PAT>....)
)/x
Note that capture buffers matched inside of recursion are not
accessible after the recursion returns, so the extra layer of
capturing buffers is necessary. Thus $+{NAME_PAT} would not
be defined even though $+{NAME} would be.
"(?>pattern)"
An "independent" subexpression, one which matches the
substring that a standalone "pattern" would match if anchored
at the given position, and it matches nothing other than this
substring. This construct is useful for optimizations of
what would otherwise be "eternal" matches, because it will
not backtrack (see "Backtracking"). It may also be useful in
places where the "grab all you can, and do not give anything
back" semantic is desirable.
For example: "^(?.)
Consider this pattern:
m{ \(
(
[^()]+ # x+
|
\( [^()]* \)
)+
\)
\)
}x
which uses "(?>...)" matches exactly when the one above does
(verifying this yourself would be a productive exercise), but
finishes in a fourth the time when used on a similar string
with 1000000 "a"s. Be aware, however, that this pattern
currently triggers a warning message under the "use warnings"
pragma or -w switch saying it "matches null string many times
in regex".
On simple groups, such as the pattern "(?> [^()]+ )", a
comparable effect may be achieved by negative look-ahead,})
Special Backtracking Control Verbs
WARNING: These patterns are experimental and subject to change or.
Verbs that take an argument
"( is optional and (*THEN) D ) /
is not the same as
/ ( A (*PRUNE) B | C (*PRUNE) D ) /
as after matching the A but failing on the B the "(*THEN)" verb
will backtrack and try C; but the "(*PRUNE)" verb will simply
fail.
"(*COMMIT)"
It is probably useful only when combined with "(?{})" or
"(??{})".
"(*ACCEPT)"
WARNING: This feature is highly experimental. It is not
recommended for production code. buffers then the
buffers are marked as ended at the point at which the
"(*ACCEPT)" was encountered. For instance:
'AB' =~ /(A (A|B(*ACCEPT)|C) D)(E)/x;
will match, and $1 will be "AB" and $2 will be "B", $3 will not
be set. If another branch in the inner parentheses were
matched, such as in the string 'ACDE', then the "D" and "E"
would have to be matched as well.. ofier. for a match. For example,
without internal optimizations done by the regular expression engine,
this will take a painfully long time to run:
'aaaaaaaaaaaa' =~ /((a{0,5}){0,5})*[c]/
And if you used "*"'s in the internal groups instead of limiting them
to 0 through 5 matches, then it would take forever--or until you ran
out of stack space. Moreover, these internal optimizations are not
always applicable. For example, if you put "{0,5}" instead of "*" on
the external group, no current optimization is applicable, and the
match takes a long time to finish.
A powerful tool for optimizing such beasts is what is known as an
"independent group", which does not backtrack (see "(?>pattern)").
Note also that zero-length look-ahead/look-behind assertions will not
backtrack to make the tail match, since they are in "logical" context:
only whether they match is considered relevant. For an example where
side-effects "\"). any character from the list. If the first
character after the "[" is "^", the class matches any character not in
the list. Within a list, the "-" character specifies a range, so that
"a-z" represents all characters between "a" and "z", inclusive. If you
want either "-" or "]" itself to be a member of a class, put it at the
that begin from and end at either alphabetics. Similarly, \xnn, where nn are hexadecimal digits, matches the
character whose numeric value is nn. The expression \cx matches the Instead of an "s///" is a double-quoted string.
with "${1}000". The operation of interpolation should not be confused
with the operation of matching a backreference. Certainly they mean
two different things on the left side of the "s///".
Repeated Patterns Matching a?" matches at the beginning of 'foo', and since the position in
the string is not moved by the match, "o?" would match again and again
because of the "*" quantifier.;
results in "<> | http://www.linux-directory.com/man1/perlre.shtml | crawl-003 | refinedweb | 2,454 | 64 |
import "android.googlesource.com/platform/tools/gpu/binary"
Package binary implements encoding and decoding of various primitive data types to and from a binary stream. The package holds BitStream for packing and unpacking sequences of bits, Float16 for dealing with 16 bit floating- point values and Reader/Writer for encoding and decoding various value types to a binary stream. There is also the higher level Encoder/Decoder that can be used for serializing object hierarchies.
binary.Reader and binary.Writer provide a symmetrical pair of methods for encoding and decoding various data types to a binary stream. For performance reasons, each data type has a separate method for encoding and decoding rather than having a single pair of methods encoding and decoding boxed values in an interface{}.
binary.Encoder and binary.Decoder extend the binary.Reader and binary.Writer interfaces by also providing a symmetrical pair of methods for encoding and decoding object types.
const IDSize = 20
IDSize is the size of an ID.
type BitStream struct { Data []byte // The byte slice containing the bits ReadPos uint32 // The current read offset from the start of the Data slice (in bits) WritePos uint32 // The current write offset from the start of the Data slice (in bits) }
BitStream provides methods for reading and writing bits to a slice of bytes. Bits are packed in a least-significant-bit to most-significant-bit order.
func (s *BitStream) Read(count uint32) uint32
Read reads the specified number of bits from the BitStream, increamenting the ReadPos by the specified number of bits and returning the bits packed into a uint32. The bits are packed into the uint32 from LSB to MSB.
func (s *BitStream) ReadBit() uint32
ReadBit reads a single bit from the BitStream, incrementing ReadPos by one.
func (s *BitStream) Write(bits, count uint32)
Write writes the specified number of bits from the packed uint32, increamenting the WritePos by the specified number of bits. The bits are read from the uint32 from LSB to MSB.
func (s *BitStream) WriteBit(bit uint32)
WriteBit writes a single bit to the BitStream, incrementing WritePos by one.
type Class interface { // ID should be a sha1 has of the types signature, such that // no two classes generate the same ID, and any change to the types name or // fields causes it's id to change. ID() ID // Encode writes the supplied object to the supplied Encoder. // The object must be a type the Class understands, the implementation is // allowed to panic if it is not. Encode(Encoder, Object) error // Decode reads a single object from the supplied Decoder. Decode(Decoder) (Object, error) // DecodeTo reads into the supplied object from the supplied Decoder. // The object must be a type the Class understands, the implementation is // allowed to panic if it is not. DecodeTo(Decoder, Object) error // Skip moves over a single object from the supplied Decoder. // This must skip the same data that Decode would have read. Skip(Decoder) error }
Class represents a struct type in the binary registry.
type Decoder interface { Reader // ID decodes a binary.ID from the stream. ID() (ID, error) // SkipID skips over a binary.ID in the stream. SkipID() error // Value decodes an Object from the stream. Value(Object) error // SkipValue must skip the same data that a call to Value would read. // The value may be a typed nil. SkipValue(Object) error // Variant decodes and returns an Object from the stream. The Class in the // stream must have been previously registered with binary.registry.Add. Variant() (Object, error) // SkipVariant must skip the same data that a call to Variant would read. SkipVariant() (ID, error) // Object decodes and returns an Object from the stream. Object instances // that were encoded multiple times may be decoded and returned as a shared, // single instance. The Class in the stream must have been previously // registered with binary.registry.Add. Object() (Object, error) // SkipObject must skip the same data that a call to Object would read. SkipObject() (ID, error) }
Decoder extends Reader with additional methods for decoding objects.
type Encoder interface { Writer // ID writes a binary.ID to the stream. ID(ID) error // Object encodes an Object with no type preamble and no sharing. Value(obj Object) error // Variant encodes an Object with no sharing. The type of obj must have // been previously registered with binary.registry.Add. Variant(obj Object) error // Object encodes an Object, optionally encoding objects only on the first // time it sees them. The type of obj must have been previously registered // with binary.registry.Add. Object(obj Object) error }
Encoder extends Writer with additional methods for encoding objects.
type Float16 uint16
Float16 represents a 16-bit floating point number, containing a single sign bit, 5 exponent bits and 10 fractional bits. This corresponds to IEEE 754-2008 binary16 (or half precision float) type.
MSB LSB ╔════╦════╤════╤════╤════╤════╦════╤════╤════╤════╤════╤════╤════╤════╤════╤════╗ ║Sign║ E₄ │ E₃ │ E₂ │ E₁ │ E₀ ║ F₉ │ F₈ │ F₇ │ F₆ │ F₅ │ F₄ │ F₃ │ F₂ │ F₁ │ F₀ ║ ╚════╩════╧════╧════╧════╧════╩════╧════╧════╧════╧════╧════╧════╧════╧════╧════╝ Where E is the exponent bits and F is the fractional bits.
func NewFloat16(f32 float32) Float16
NewFloat16 returns a Float16 encoding of a 32-bit floating point number. Infinities and NaNs are encoded as such. Very large and very small numbers get rounded to infinity and zero respectively.
func NewFloat16Inf(sign int) Float16
Float16Inf returns positive infinity if sign >= 0, negative infinity if sign < 0.
func NewFloat16NaN() Float16
Float16NaN returns an “not-a-number” value.
func (f Float16) Float32() float32
Float32 returns the Float16 value expanded to a float32. Infinities and NaNs are expanded as such.
func (f Float16) IsInf(sign int) bool
IsInf reports whether f is an infinity, according to sign. If sign > 0, IsInf reports whether f is positive infinity. If sign < 0, IsInf reports whether f is negative infinity. If sign == 0, IsInf reports whether f is either infinity.
func (f Float16) IsNaN() bool
IsNaN reports whether f is an “not-a-number” value.
type Generate struct{}
Generate is used to tag structures that need an auto generated Class. The codergen function searches packages for structs that have this type as an anonymous field, and then automatically generates the encoding and decoding functionality for those structures. For example, the following struct would create the Class with methods needed to encode and decode the Name and Value fields, and register that class. The embedding will also fully implement the binary.Object interface, but with methods that panic. This will get overridden with the generated Methods. This is important because it means the package is resolvable without the generated code, which means the types can be correctly evaluated during the generation process.
type MyNamedValue struct {
binary.Generate Name string Value []byte
}
func (Generate) Class() Class
type ID [IDSize]byte
ID is a codeable unique identifier.
func NewID(data ...[]byte) ID
Create a new ID that is the sha1 hash of the supplied data.
func ParseID(s string) (id ID, err error)
ParseID parses lowercase string s as a 20 byte hex-encoded ID.
func (id ID) Format(f fmt.State, c rune)
func (id ID) String() string
func (id ID) Valid() bool
Valid returns true if the id is not the default value.
type Object interface { // Class returns the serialize information and functionality for this type. // The method should be valid on a nil pointer. Class() Class }
Object is the interface to any class that wants to be encoded/decoded.
type Reader interface { // Data writes the data bytes in their entirety. Data([]byte) error // Skip jumps past count bytes. Skip(count uint32) error // Bool decodes and returns a boolean value from the Reader. Bool() (bool, error) // Int8 decodes and returns a signed, 8 bit integer value from the Reader. Int8() (int8, error) // Uint8 decodes and returns an unsigned, 8 bit integer value from the Reader. Uint8() (uint8, error) // Int16 decodes and returns a signed, 16 bit integer value from the Reader. Int16() (int16, error) // Uint16 decodes and returns an unsigned, 16 bit integer value from the Reader. Uint16() (uint16, error) // Int32 decodes and returns a signed, 32 bit integer value from the Reader. Int32() (int32, error) // Uint32 decodes and returns an unsigned, 32 bit integer value from the Reader. Uint32() (uint32, error) // Float32 decodes and returns a 32 bit floating-point value from the Reader. Float32() (float32, error) // Int64 decodes and returns a signed, 64 bit integer value from the Reader. Int64() (int64, error) // Uint64 decodes and returns an unsigned, 64 bit integer value from the Reader. Uint64() (uint64, error) // Float64 decodes and returns a 64 bit floating-point value from the Reader. Float64() (float64, error) // String decodes and returns a string from the Reader. String() (string, error) // SkipString skips over a single string from the Reader. SkipString() error }
Reader provides methods for decoding values.
type Writer interface { // Data writes the data bytes in their entirety. Data([]byte) error // Bool encodes a boolean value to the Writer. Bool(bool) error // Int8 encodes a signed, 8 bit integer value to the Writer. Int8(int8) error // Uint8 encodes an unsigned, 8 bit integer value to the Writer. Uint8(uint8) error // Int16 encodes a signed, 16 bit integer value to the Writer. Int16(int16) error // Uint16 encodes an unsigned, 16 bit integer value to the Writer. Uint16(uint16) error // Int32 encodes a signed, 32 bit integer value to the Writer. Int32(int32) error // Uint32 encodes an usigned, 32 bit integer value to the Writer. Uint32(uint32) error // Float32 encodes a 32 bit floating-point value to the Writer. Float32(float32) error // Int64 encodes a signed, 64 bit integer value to the Writer. Int64(int64) error // Uint64 encodes an unsigned, 64 bit integer value to the Encoders's io.Writer. Uint64(uint64) error // Float64 encodes a 64 bit floating-point value to the Writer. Float64(float64) error // String encodes a string to the Writer. String(string) error }
Writer provides methods for encoding values. | https://android.googlesource.com/platform/tools/gpu/+/refs/heads/studio-1.3-release/binary/ | CC-MAIN-2022-27 | refinedweb | 1,628 | 57.47 |
Let's start with the new bit of code you saw in exercise 1:
def alphabetize(arr, rev=false)
The first part makes sense—we're defining a method,
alphabetize. We can guess that the first parameter is an array, but what's this
rev=false business?
What this does is tell Ruby that
alphabetize has a second parameter,
rev (for "reverse") that will default to
false if the user doesn't type in two arguments. You might have noticed that our first call to
alphabetize in exercise 1 was just
alphabetize(books)
Ruby didn't see a
rev, so it gave it the default value of
false. | https://www.codecademy.com/courses/learn-ruby/lessons/ordering-your-library/exercises/default-parameters | CC-MAIN-2018-39 | refinedweb | 108 | 63.53 |
On Wed, 8 Aug 2001, Ralf Baechle wrote:
> On Tue, Aug 07, 2001 at 08:45:17PM -0400, J. Scott Kasten wrote:
> Right circumstances = almost always. Actually Linux makes sure that it's
> indeed always.
As long as our data doesn't make sparse use of the VM address space,
yes... I always have to think of the worst case. ;-)
> > Some of the old hands here could tell you better how Irix behaves on those
> > boxes. I know you can compile code with 64 bit int and pointers and it
> > will run on those boxes under Irix, but there is a little more to it than
> > that.
>
> That's not supported on all systems. An Indy for example is limited to
> 128mb RAM as shipped by SGI (256 with aftermarket parts). There is no
> point in supporting 64-bit address space on such a system and so SGI doesn't
> support only N32.
Well, that actually brings to mind a nagging question I've had in regards
to just what N32 is. Here's an example app and creative use of gcc:
/* Note: stdio.h deliberately left out to bypass some syntax issues. */
int main(int argc, char *agv[]) {
int i, *j;
printf(
"sizeof(int) = %1d, sizeof(*) = %1d\n",
sizeof(i),
sizeof(j)
);
j = &i;
*j = 10;
i++;
printf(
"Result: %1d\n",
(int) *j
);
return (0);
}
Compile:
gcc -mips3 -mint64 test.c -o test
The file command says:
test: ELF N32 MSB mips-3 dynamic executable (not stripped) MIPS -
version 1
Run:
sizeof(int) = 8, sizeof(*) = 8
Result: 11
If we look at the assembly, we see a sign extended 64 bit load, and a 64
bit add. So we are indeed generating 64 bit instructions, at least in
some cases.
dli $3,0xa # 10
<snip>
daddu $3,$2,1
Does N32 legitimately allow 64 bit instructions, or is this an example of
code that I've truely "munged" togeather? It clearly works, at least in
this trivial case.
I'm just trying to solidify my understanding of what Irix is supposed to
do. | https://www.linux-mips.org/archives/linux-mips/2001-08/msg00105.html | CC-MAIN-2016-36 | refinedweb | 344 | 71.34 |
Introduction
Dependencies are essential for modern development. They save you time and energy. Functionalities you may need for your app like sending e-mails or logging can all be easily included as third party libraries. Thanks to the open source movement, there are many high quality packages to choose from. In the early days, including third-party libraries was cumbersome and error prone but luckily, today we have tools like Composer to help us.
Composer is an exceptional dependency manager for PHP. It replaces PEAR and rightfully so. PEAR requires that your project is specially prepared to work with it, where Composer gives you all the freedom you need without any special requirements. One major difference between these two tools is that PEAR installs dependencies globally and Composer installs them locally, in your project structure. PEAR is essentially a package manager and Composer is a dependency manager. That’s where the emphasis is.
Composer was inspired by projects like NPM and Bundler. The vast selection of compatible packages are hosted on the official Composer repository called Packagist. These packages are open source so you can contribute to them too. Popular frameworks and tools like Laravel, PHPUnit and Monolog can all be found here. You can even use a specific code revision of the package when including it in your project so you are getting great flexibility. Composer packages are versioned, so you can pin down the exact version of the package you need. This makes porting your project to another machine or to a CI service such as [Semaphore] () effortless.
In this tutorial, we’ll explore some of the most used Composer features and show how to use them. After following through, you should be comfortable with managing your PHP project’s dependencies with Composer.
Prerequisites
The software you need is as follows:
- Some version of PHP 5, preferably the latest. Composer is compatible with PHP versions 5.3.2 and up.
- Clients for Git and Subversion
I’ll be using PHP version 5.6.5, but it’s perfectly fine to use any other version you have as long as it’s newer than 5.3.2. If you don’t have PHP installed, you can do it with phpbrew which makes managing multiple PHP versions on the same machine easy.
Installation
Composer can be installed in two different ways.
Install Locally
Local installation will download
composer.phar to the current directory. The drawback of this method is that you’ll always have to reference the Composer’s executable from the directory where it’s downloaded to.
$ curl -sS | php All settings correct for using Composer Downloading... Composer successfully installed to: /workspace/composer.phar Use it: php composer.phar $ php composer.phar --version Composer version 1.0-dev (1d8f05f1dd0e390f253f79ea86cd505178360019)
Install Globally (Recommended)
Installing Composer globally is a handy way to have access to the tool from anywhere by just executing the
composer command.
$ curl -sS | php All settings correct for using Composer Downloading... Composer successfully installed to: /workspace/composer.phar Use it: php composer.phar $ sudo mv composer.phar /usr/local/bin/composer $ composer --version Composer version 1.0-dev (1d8f05f1dd0e390f253f79ea86cd505178360019)
If you installed PHP with
phpbrew running the command below is sufficient
phpbrew install-composer
We continue this tutorial with the assumption that Composer has been installed globally.
Configuring Composer
Composer is configured with a single file named
composer.json located in the root directory of the project. To follow along, just create an empty directory called composer-tutorial, fire up a text editor and create an empty
composer.json file in this directory. Here’s how the most simple Composer file looks like:
{ "require": { "symfony/yaml": "2.6.4" } }
The only requirement other than the fact that the file has to be in a JSON format, is the inclusion of the
require key. This key defines your project’s dependencies with a list of packages and their respective versions.
Defining Dependencies
By convention, package names consist of the package’s vendor and its project name. This is done in an effort to avoid name conflicts. In our example above, the vendor is
symfony and the project is called
yaml.
The version of the package can be defined in multiple ways. This is where Composer gives you huge flexibility.
2.6.4, locks the package to the exact version defined
>= 2.6.4, defines a range where the package’s version has to be at least
2.6.4but a newer version is used if it’s available. Operators like
<,
<=,
>,
>=and
!=can also be used.
2.6.*, which is equivalent to
>=2.6 <2.7.
~2.6, is used for projects which use semantic versioning and is equivalent to
>=2.6 <3.0. It essentially allows the micro version to change:
~2.6.4is the same as saying
>=2.6.4 <2.7.0.
^2.6, caret is used when we want to avoid braking updates for projects which employ semantic versioning.
^2.6is equivalent to
>=2.6.4 <3.0.0.
A more detailed description of these restrictions can be found on [this page] ().
Breaking Out of the Defaults
If you’re not satisfied with the default settings, you can use the
config key and customize anything from specifying Composer’s default installation path to defining GitHub protocols to use. For example, to define a custom installation path for our Composer packages, we alter our
composer.json to look like this:
{ "config": { "vendor-dir": "dependencies" }, "require": { "symfony/yaml": "2.6.4" } }
A full list of
config options can be found here.
Custom package sources
A package is essentially a directory containing information like the version of the package and the source from where to get the contents of the package. Two types of package sources exist:
dist, a packaged version, tested and stable
source, clones the contents of the package from a version control system
Packages from
source are usually used in development and
dist packages tend to be production ready.
Composer by default uses
dist packages from Packagist. However, this can be easily expanded by using the
repositories key and including
source packages. You can use sources like version control systems (Git, SVN and Hg) or even PEAR packages. To add the
monolog library directly from GitHub, the following has to be done:
{ "repositories": [ { "type": "vcs", "url": "" } ], "require": { "symfony/yaml": "2.6.4", "monolog/monolog": "1.12.0" } }
Repositories are a list of versions for the package. We require a version from the repository the same way we do with
dist packages.
Refer to this page for an in-depth explanation of the various options for configuring repositories.
Installing dependencies
To install the dependencies you defined in
composer.json, we use the following command:
composer install
The installation process will fetch the latest packages according to the constraints we defined. The packages are put in the
vendor directory or to a custom directory like we defined earlier. The installation command has a couple of interesting options.
--prefer-sourcewill install the packages from their source which is usually a GitHub or a Subversion repository. In other words, it clones the package’s source. In the case where the repository is not found on the vcs, it falls back to the installation from
dist.
--prefer-distprefers to install packages from Packagist and caches the package archives locally.
--profilewill give you statistics about the installation at the end of the process. It can be used with any Composer command.
Caching dependencies
Composer stores its cache in
$COMPOSER_HOME/cache which is usually set to
~/.composer/cache by default. Caching is very beneficial when there are multiple projects which have some dependencies in common. The first time
composer install is executed, it downloads and caches the dependencies which are not present in the cache. The next time it’s initiated, it’ll load the installed packages from cache.
Let’s see what difference does the installation from cache can make.
user@host:~/workspace/project-1/$ composer install --prefer-dist --profile [6.5MB/0.03s] Loading composer repositories with package information [6.8MB/0.48s] Installing dependencies (including require-dev) [47.0MB/12.45s] - Installing symfony/yaml (v2.6.4) [47.1MB/13.70s] Downloading: 0%[47.1MB/13.71s] ... [47.1MB/13.93s] Downloading: 100%[47.1MB/13.93s] [46.8MB/14.37s] Writing lock file [46.8MB/14.37s] Generating autoload files [46.8MB/14.37s] Memory usage: 46.77MB (peak: 56.15MB), time: 14.37s
Here’s what happens when a cached packaged is reused for another installation:
user@host:~/workspace/project-2/$ composer install --prefer-dist --profile [6.1MB/0.03s] Loading composer repositories with package information [6.4MB/0.41s] Installing dependencies (including require-dev) [46.5MB/3.65s] - Installing symfony/yaml (v2.6.4) [46.9MB/3.65s] Loading from cache [47.0MB/3.68s] [46.6MB/4.07s] Writing lock file [46.6MB/4.07s] Generating autoload files [46.6MB/4.07s] Memory usage: 46.62MB (peak: 55.65MB), time: 4.07s
As you can see, installing cached packages is much faster the second time. This makes a big difference in a project with many dependencies.
Locking down dependencies
When the installation process is initiated the first time, it creates a
composer.lock file. This file locks down all the currently installed packages with their versions and dependencies. It’s important to include this file to the version control system, because this is what makes your project transferable to other machines. This way you can rest assured that anyone who contributes to the project will have the same dependencies installed as you.
Updating dependencies
Frameworks and libraries are being improved all the time and if we want to get the latest features and bugfixes, our dependencies need to be kept up to date. This can be achieved simply by using the
update command.
composer update
The
update command will install all the latest packages while taking into consideration the version constraints defined in
composer.json and it also updates the
composer.lock file.
Updating all the dependencies however isn’t always optimal. Maybe there’s a breaking change in one of the updates and you want to keep the old version until you handle it later. Composer packages can be updated one-by-one by letting the
update command which
vendor/project you want updated.
composer update symfony/yaml
The update command has very similar options to
install and they can be found here.
During your time with installing and updating Composer packages, you might meet this error message:
Warning: The lock file is not up to date with the latest changes in composer.json, you may be getting outdated dependencies, run update to update them.
This happens when even a tiniest adjustment is introduced in
composer.json. Like changing the author, description or even altering a single letter. The slightest modification is enough to change the file’s MD5 hash which is detected by
composer.lock. To avoid running the
update when a trivial modification occurs, the
--lock option is used to suppress the warning and update the lock file.
composer update --lock
Autoloading
You may have noticed that Composer is generating autoload files after the packages are successfully installed or updated. More specifically, the file in question is
vendor/autoload.php. Autoloading makes it really convenient to use the available dependencies throughout your project by simply referencing their class names without the need to explicitly require the files where the classes reside. The only thing that needs to be included in your project is the autoload file.
Composer currently supports these autoloading mechanisms:
There’s even an option to autoload your own classes which we will demonstrate next.
Autoloading a custom class
Let’s see a simple example about how autoloading works. In our root directory
composer-tutorial, create a directory called
source. In this newly created directory, create a file called
Square.php.
<?php namespace Shapes; class Square { static function area($side){ $surface_area = $side * $side; echo "The squares surface area is $surface_area units\n"; } }
The
area function calculates the square’s surface area according to its side.
Add the
autoload key to
composer.json where we define the mapping between the namespace and the path. Composer will look for files located in
./source which belongs to the
/Shapes namespace.
// composer.json { "autoload": { "psr-4": {"Shapes\\": "source/"} }, "repositories": [ { "type": "vcs", "url": "" } ], "require": { "symfony/yaml": "2.6.4" } }
After the custom class is included the autoload file has to be regenerated to reflect the changes.
composer dump-autoload
You can see the effect of
dump-autoload in
vendor/composer/autoload_psr4.php.
<?php // autoload_psr4.php @generated by Composer $vendorDir = dirname(dirname(__FILE__)); $baseDir = dirname($vendorDir); return array( 'Shapes\\' => array($baseDir . '/source'),' )
Now, let’s create a simple command line script which will use the
Square class by just requiring
vendor/autoload.php and nothing more. Create this file in the root directory (
composer-tutorial) and name it
calculate_area.php.
#! /usr/bin/env php <?php require_once __DIR__ . '/vendor/autoload.php'; $side = isset($argv[1]) ? $argv[1] : null; if (isset($side) && ctype_digit($side)) { Shapes\Square::area($side); } else { echo "Please provide a valid size!\n"; }
After the file is created, make it executable and run it.
$ chmod +x ./calculate_area.php $ ./calculate_area.php 40 The square's surface area is 1600 units
Our
Shapes\Square class was successfully loaded without the need to explicitly include it.
Wrapping it up
In this tutorial we saw how Composer manages dependencies for PHP projects. This versatile tool is now part of many developers’ arsenal. I hope this article managed to bring you up to speed with Composer.
For a detailed overview of the capabilities of the command line interface, refer to the Composer documentation. | https://semaphoreci.com/community/tutorials/getting-started-with-composer-for-php-dependency-management | CC-MAIN-2019-47 | refinedweb | 2,283 | 58.99 |
[
]
Konstantin Shvachko commented on HDFS-5453:
-------------------------------------------
It is indeed strange as Suresh mentioned, that get block locations performance in ASYNC falls
way below BASE with 15 and 16 threads. Not asking what happens with 200 threads only because
you refer to it as a simulation with a radically different architecture, which should probably
be discussed elsewhere.
BTW, for the BASE case you should see improvement with 200 handlers as it starts benefiting
from edits batching.
As for fine grain locking it looks to me that your main motivation is not the performance
but rather the new functionality pursued with HDFS-5477. BTW should it be linked here then.
I see you propose to separate BM and access it with rpc calls from FSNamesystem. But I don't
quite understand the consistency model. With fine or coarse locking you will hold a lock while
making an rpc call. If the rpc fails you will need to provide consistency of the namespace
state and the retries with e.g. at-most-once semantics, mentioned in the design. So why not
just release the lock while making the rpc and reacquire it when the rpc is completed. In
a sense the lock here is a flag signalling a BM call is in progress so that others could avoid
accessing the INode.
I mean one way or another a consistency logic should be in place to handle failures. And may
be there is a way to design the feature without relying on locking to provide consistency
in general.
> Support fine grain locking in FSNamesystem
> ------------------------------------------
>
> Key: HDFS-5453
> URL:
> Project: Hadoop HDFS
> Issue Type: New Feature
> Components: namenode
> Affects Versions: 2.0.0-alpha, 3.0.0
> Reporter: Daryn Sharp
> Assignee: Daryn Sharp
> Attachments: async_simulation.xlsx
>
>
>.5#6160) | http://mail-archives.apache.org/mod_mbox/hadoop-hdfs-issues/201312.mbox/%3CJIRA.12677391.1383580476930.20806.1388362310486@arcas%3E | CC-MAIN-2018-30 | refinedweb | 293 | 62.27 |
I have a training project at my new job and a really small but crucial
setback.
I have a datagridview control bound to a dataset read
from an xml file. It has 3 columns: item, price, calories. I am supposed to
calculate the average values for price and calories and place them in the
row below the entries that are read in from the xml. This works fine,
except that when I fill the bo
When calling the Add method in the Appointments controller a new entry
is created in the student table even when there is a matching student.
What I want is to update the matching student entry based on
the foreign key which has already been established. Of course if there is
no match then it should insert.
public class Appointment{ public int Id { get
I have a simple Crud application where there is one filed called
Category. During Creation , for Category I have provided a simple drop down
box which lists all categories. During Editing, I need to get the same drop
down box with the entry in the database chosen.How do i do this.
For creation i used say
<p><b>Category:</b> &nb
I have a document library site and would like to send an email when a
document object is edited, containing a summary of the changes.
The database interaction is Code First Entities Framework using
DBContext
Here is what I have so far:
[HttpPost] public ActionResult Edit(Document document, bool
sendEmail, string commentsTextBox) {
Recording
changed
values
with
DBContext
Entry
That is a hard one I searched everywhere and could find a resobale
answer. What I want is:
1) Read the value "ServiceName" value
for each entry in "HKEY_LOCAL_MACHINESOFTWAREMicrosoftWindows
NTCurrentVersionNetworkCards"
2) Create a new Registry for each
"ServiceName" from each "Entry" on
"HKEY_LOCAL_MACHINESYSTEMControlSet001ControlClass{4D36E972-E325-11CE-BFC1-08002BE1
Trying to Post Messages to IBM Connections using http POST method using
c#.
ATOM formatted Message Entry Message XML:
<entry xmlns="">
<category scheme=""
term="entry"/> <category
scheme=""
term="simpleEntry"/> <content
I was once told that it is faster to just run an insert and let the
insert fail than to check if a database entry exists and then inserting if
it is missing.
I was also told that that most databases are
heavily optimized for reading reading rather than writing, so wouldn't a
quick check be faster than a slow insert?
Is this a question of
the expected number of collisions
I am using the TrueZip library to inject an entry to a .tar.gz file.
On running my code, I find the file I want to inject getting pushed to
the .tar.gz file. But the inserted entry has zero size, i.e. it appears
that only a placeholder gets created for that entry and its contents are
not copied. Here is the piece of code:
public class
FileCompressUtil {private s
I am trying to create a search from ProductDB(database), the main
columns I would like the user to search is Material_No and
Product_Line.
So far, I have the following:
Drop
Down List:
<asp:DropDownList
I have a Product table with a field to store the price. Each time,
this price was changing, i need te keep the ancient value. So I have
created a table called History to store the product id, the price and the
date of changing.I've also created this method in my product class to
generate a new entry :
public function
updateHistory($modified) { if(array_ | http://bighow.org/tags/entry/1 | CC-MAIN-2017-04 | refinedweb | 580 | 57.3 |
Chex: Testing made fun, in JAX!
Project description
Chex
Chex is a library of utilities for helping to write reliable JAX code.
This includes utils to help:
- Instrument your code (e.g. assertions)
- Debug (e.g. transforming
pmapsin
vmapswithin a context manager).
- Test JAX code across many
variants(e.g. jitted vs non-jitted).
Installation
Chex can be installed with pip directly from github, with the following command:
pip install git+git://github.com/deepmind/chex.git
or from PyPI:
pip install chex
Modules Overview
Dataclass (dataclass.py)
Dataclasses are a popular construct introduced by Python 3.7 to allow to easily specify typed data structures with minimal boilerplate code. They are not, however, compatible with JAX and dm-tree out of the box.
In Chex we provide a JAX-friendly dataclass implementation reusing python dataclasses.
Chex implementation of
dataclass registers dataclasses as internal PyTree
nodes to ensure
compatibility with JAX data structures.
In addition, we provide a class wrapper that exposes dataclasses as
collections.Mapping descendants which allows to process them
(e.g. (un-)flatten) in
dm-tree methods as usual Python dictionaries.
See
@mappable_dataclass
docstring for more details.
Example:
@chex.dataclass class Parameters: x: chex.ArrayDevice y: chex.ArrayDevice parameters = Parameters( x=jnp.ones((2, 2)), y=jnp.ones((1, 2)), ) # Dataclasses can be treated as JAX pytrees jax.tree_map(lambda x: 2.0 * x, parameters) # and as mappings by dm-tree tree.flatten(parameters)
NOTE: Unlike standard Python 3.7 dataclasses, Chex
dataclasses cannot be constructed using positional arguments. They support
construction arguments provided in the same format as the Python dict
constructor. Dataclasses can be converted to tuples with the
from_tuple and
to_tuple methods if necessary.
parameters = Parameters( jnp.ones((2, 2)), jnp.ones((1, 2)), ) # ValueError: Mappable dataclass constructor doesn't support positional args.
Assertions (asserts.py)
One limitation of PyType annotations for JAX is that they do not support the
specification of
DeviceArray ranks, shapes or dtypes. Chex includes a number
of functions that allow flexible and concise specification of these properties.
E.g. suppose you want to ensure that all tensors
t1,
t2,
t3 have the same
shape, and that tensors
t4,
t5 have rank
2 and (
3 or
4), respectively.
chex.assert_equal_shape([t1, t2, t3]) chex.assert_rank([t4, t5], [2, {3, 4}])
More examples:
from chex import assert_shape, assert_rank, ... assert_shape(x, (2, 3)) # x has shape (2, 3) assert_shape([x, y], [(), (2,3)]) # x is scalar and y has shape (2, 3) assert_rank(x, 0) # x is scalar assert_rank([x, y], [0, 2]) # x is scalar and y is a rank-2 array assert_rank([x, y], {0, 2}) # x and y are scalar OR rank-2 arrays assert_type(x, int) # x has type `int` (x can be an array) assert_type([x, y], [int, float]) # x has type `int` and y has type `float` assert_equal_shape([x, y, z]) # x, y, and z have equal shapes assert_tree_all_close(tree_x, tree_y) # values and structure of trees match assert_tree_all_finite(tree_x) # all tree_x leaves are finite assert_devices_available(2, 'gpu') # 2 GPUs available assert_tpu_available() # at least 1 TPU available assert_numerical_grads(f, (x, y), j) # f^{(j)}(x, y) matches numerical grads
JAX re-traces JIT'ted function every time the structure of passed arguments
changes. Often this behavior is inadvertent and leads to a significant
performance drop which is hard to debug. @chex.assert_max_traces
decorator asserts that the function is not re-traced more that
n times during
program execution.
Global trace counter can be cleared by calling
chex.clear_trace_counter(). This function be used to isolate unittests relying
on
@chex.assert_max_traces.
Examples:
@jax.jit @chex.assert_max_traces(n=1) def fn_sum_jitted(x, y): return x + y z = fn_sum_jitted(jnp.zeros(3), jnp.zeros(3)) t = fn_sum_jitted(jnp.zeros(6, 7), jnp.zeros(6, 7)) # AssertionError!
Can be used with
jax.pmap() as well:
def fn_sub(x, y): return x - y fn_sub_pmapped = jax.pmap(chex.assert_max_retraces(fn_sub), n=10)
See documentation of asserts.py for details on all supported assertions.
Test variants (variants.py)
JAX relies extensively on code transformation and compilation, meaning that it can be hard to ensure that code is properly tested. For instance, just testing a python function using JAX code will not cover the actual code path that is executed when jitted, and that path will also differ whether the code is jitted for CPU, GPU, or TPU. This has been a source of obscure and hard to catch bugs where XLA changes would lead to undesirable behaviours that however only manifest in one specific code transformation.
Variants make it easy to ensure that unit tests cover different ‘variations’ of a function, by providing a simple decorator that can be used to repeat any test under all (or a subset) of the relevant code transformations.
E.g. suppose you want to test the output of a function
fn with or without jit.
You can use
chex.variants to run the test with both the jitted and non-jitted
version of the function by simply decorating a test method with
@chex.variants, and then using
self.variant(fn) in place of
fn in the body
of the test.
def fn(x, y): return x + y ... class ExampleTest(chex.TestCase): @chex.variants(with_jit=True, without_jit=True) def test(self): var_fn = self.variant(fn) self.assertEqual(fn(1, 2), 3) self.assertEqual(var_fn(1, 2), fn(1, 2))
If you define the function in the test method, you may also use
self.variant
as a decorator in the function definition. For example:
class ExampleTest(chex.TestCase): @chex.variants(with_jit=True, without_jit=True) def test(self): @self.variant def var_fn(x, y): return x + y self.assertEqual(var_fn(1, 2), 3)
Example of parameterized test:
from absl.testing import parameterized # Could also be: # `class ExampleParameterizedTest(chex.TestCase, parameterized.TestCase):` # `class ExampleParameterizedTest(chex.TestCase):` class ExampleParameterizedTest(parameterized.TestCase): @chex.variants(with_jit=True, without_jit=True) @parameterized.named_parameters( ('case_positive', 1, 2, 3), ('case_negative', -1, -2, -3), ) def test(self, arg_1, arg_2, expected): @self.variant def var_fn(x, y): return x + y self.assertEqual(var_fn(arg_1, arg_2), expected)
Chex currently supports the following variants:
with_jit-- applies
jax.jit()transformation to the function.
without_jit-- uses the function as is, i.e. identity transformation.
with_device-- places all arguments (except specified in
ignore_argnumsargument) into device memory before applying the function.
without_device-- places all arguments in RAM before applying the function.
with_pmap-- applies
jax.pmap()transformation to the function (see notes below).
See documentation in variants.py for more details on the supported variants. More examples can be found in variants_test.py.
Variants notes
Test classes that use
@chex.variantsmust inherit from
chex.TestCase(or any other base class that unrolls tests generators within
TestCase, e.g.
absl.testing.parameterized.TestCase).
[
jax.vmap] All variants can be applied to a vmapped function; please see an example in variants_test.py (
test_vmapped_fn_named_paramsand
test_pmap_vmapped_fn).
[
@chex.all_variants] You can get all supported variants by using the decorator
@chex.all_variants.
[
with_pmapvariant]
jax.pmap(fn)(doc) performs parallel map of
fnonto multiple devices. Since most tests run in a single-device environment (i.e. having access to a single CPU or GPU), in which case
jax.pmapis a functional equivalent to
jax.jit,
with_pmapvariant is skipped by default (although it works fine with a single device). Below we describe a way to properly test
fnif it is supposed to be used in multi-device environments (TPUs or multiple CPUs/GPUs). To disable skipping
with_pmapvariants in case of a single device, add
--chex_skip_pmap_variant_if_single_device=falseto your test command.
Fakes (fake.py)
Debugging in JAX is made more difficult by code transformations such as
jit
and
pmap, which introduce optimizations that make code hard to inspect and
trace. It can also be difficult to disable those transformations during
debugging as they can be called at several places in the underlying
code. Chex provides tools to globally replace
jax.jit with a no-op
transformation and
jax.pmap with a (non-parallel)
jax.vmap, in order to more
easily debug code in a single-device context.
For example, you can use Chex to fake
pmap and have it replaced with a
vmap.
This can be achieved by wrapping your code with a context manager:
with chex.fake_pmap(): @jax.pmap def fn(inputs): ... # Function will be vmapped over inputs fn(inputs)
The same functionality can also be invoked with
start and
stop:
fake_pmap = chex.fake_pmap() fake_pmap.start() ... your jax code ... fake_pmap.stop()
In addition, you can fake a real multi-device test environment with a multi-threaded CPU. See section Faking multi-device test environments for more details.
See documentation in fake.py and examples in fake_test.py for more details.
Faking multi-device test environments
In situations where you do not have easy access to multiple devices, you can still test parallel computation using single-device multi-threading.
In particular, one can force XLA to use a single CPU's threads as separate devices, i.e. to fake a real multi-device environment with a multi-threaded one. These two options are theoretically equivalent from XLA perspective because they expose the same interface and use identical abstractions.
Chex has a flag
chex_n_cpu_devices that specifies a number of CPU threads to
use as XLA devices.
To set up a multi-threaded XLA environment for
absl tests, define
setUpModule function in your test module:
def setUpModule(): chex.set_n_cpu_devices()
Now you can launch your test with
python test.py --chex_n_cpu_devices=N to run
it in multi-device regime. Note that all tests within a module will have an
access to
N devices.
More examples can be found in variants_test.py, fake_test.py and fake_set_n_cpu_devices_test.py.
Citing Chex
To cite this repository:
@software{chex2020github, author = {David Budden and Matteo Hessel and Iurii Kemaev and Stephen Spencer and Fabio Viola}, title = {Chex: Testing made fun, in JAX!}, url = {}, version = {0.0.1}, year = {2020}, }
In this bibtex entry, the version number is intended to be from chex/__init__.py, and the year corresponds to the project's open-source release.
Project details
Download files
Download the file for your platform. If you're not sure which to choose, learn more about installing packages. | https://pypi.org/project/chex/ | CC-MAIN-2021-10 | refinedweb | 1,680 | 51.65 |
Subject: Re: [OMPI users] ptrdiff_t undefined error on intel 64bit machine with intel compilers
From: Tamara Rogers (talmesha_r_at_[hidden])
Date: 2009-02-23 18:07:32
Jeff:
That patch didn't work. However, I hacked the configure file and got it to work by forcing OMPI_PTRDIFF_TYPE to be long:
line 65040 of configure
#define OMPI_PTRDIFF_TYPE long
It compiled just fine after this.
Thanks for all your efforts
---, 6:18 PM
Does applying the following patch fix the problem?
Index: ompi/datatype/dt_args.c
===================================================================
--- ompi/datatype/dt_args.c (revision 20616)
+++ ompi/datatype/dt_args.c (working copy)
@@ -18,6 +18,9 @@
*/
#include "ompi_config.h"
+
+#include <malloc.h>
+
#include "opal/util/arch.h"
#include "opal/include/opal/align.h"
#include "ompi/constants.h"
On Feb 20, 2009, at 4:33 PM, Tamara Rogers wrote:
> Jeff:
> See attached.I'm using the 9.0 version of the intel compilers.
Interestngly I have no problems on a 32bit intel machine using these same
compilers. There only seems to be a problem on the 64bit machine.
>
> ---, 8:37 AM
>
> Can you also send a copy of your mpi.h? (OMPI's mpi.h is generated by
> configure; I want to see what was put into your mpi.h)
>
> Finally, what version of icc are you using? I test regularly with icc
9.0,
> 9.1, 10.0, and 10.1 with no problems. Are you using newer or older? (I
> don't have immediate access to 11.x or 8.x)
>
>
> On Feb 20, 2009, at 8:09 AM, Jeff Squyres wrote:
>
> > Can you send your config.log as well?
> >
> > It looks like you forgot to specify FC=ifort on your configure line
(i.e.,
> you need to specify F77=ifort for the Fortran 77 *and* FC=ifort for the
Fortran
> 90 compiler -- this is an Autoconf thing; we didn't make it up).
> >
> > That shouldn't be the problem here, but I thought I'd mention
it.
> >
> >
> > On Feb 19, 2009, at 12:00 PM, Tamara Rogers wrote:
> >
> >>
> >>
> >>
> >>
> >>
> >>
>
<openmp-1.3_output.tar.gz>_______________________________________________
> >> users mailing list
> >> users_at_[hidden]
> >>
> >
> >
> > --Jeff Squyres
> > Cisco Systems
> >
> > _______________________________________________
> > users mailing list
> > users_at_[hidden]
> >
>
>
> --Jeff Squyres
> Cisco Systems
>
> _______________________________________________
> users mailing list
> users_at_[hidden]
>
>
>
<openmpi-1.3_64_output.tar.gz>_______________________________________________
> users mailing list
> users_at_[hidden]
>
--Jeff Squyres
Cisco Systems
_______________________________________________
users mailing list
users_at_[hidden] | http://www.open-mpi.org/community/lists/users/2009/02/8170.php | CC-MAIN-2014-52 | refinedweb | 385 | 69.58 |
Before talking about the code, there are a couple of things that I want to point out. One, this is my first article and I'm a little bit excited. Two, I'm dealing with software development for personal purposes in an amateur fashion. They are probably not the best-practices. But my purpose is to reflect my ideas, find solution to my problems. Three, I have been following CodeProject for the past 1 year and I have learnt lots of things here, as a result I cannot distinguish from whom I have learnt something. So, if you think that some of the ideas which are presented here are yours, please contact me so that I can refer it to you.
Now, let's talk about the project in brief. This is mainly a server-client chat application designed for network environment. I have modified it so that it can be run on a terminal server/terminal client environment. I mean users log on a terminal server with a remote desktop connection client (Microsoft remote desktop connection, hoblink, etc.). I chose to make the server a console application to track users easily (who logged on or off...). But there are better solutions that can be implemented for the server-side.
There are many chat solutions available here on CodeProject and also on the internet. I have tried to implement them, but in one way or the other they all seem to be a little bit confusing. At last I managed to find a solution as a result of a lot of digging. The solution not only consists of server side and client side applications, but also a linked list solution to keep users' information on the server-side. In fact, the linked list solution can be the subject of another article by itself, but I don't have that much time, so I will try to explain that here in brief. I have tried to give plenty of comments in the code to guide the users.
The solution consists of five projects:
ServerApp
ServerClass
ClientApp
ClientClass
LinkedList
I made this separation for the sake of reusability.
ServerApp is a console application.
ServerApp
namespace ServerAppSpace //Namespace for the Server Application
{
//IServerApp is the interface for the server application to implement...
//it will provide necessary methods to interact with the server
class MsgServer : MarshalByRefObject, IServerApp
{
#region IServerApp implementation
...
#endregion
[STAThread]
public static void Main(string[] args)
{
TcpChannel channel = new TcpChannel(9001);
ChannelServices.RegisterChannel(channel);
ServerClass remService = new ServerClass();
ObjRef obj = RemotingServices.Marshal(remService,"TcpService");
// Create applications MainForm
MsgServer frmMain = new MsgServer();
// provide marshaled object with reference to Application
remService.theMainServer = ( IServerApp) frmMain;
...
RemotingServices.Unmarshal(obj);
RemotingServices.Disconnect(remService);
}
}
ServerClass holds the necessary interfaces (IServerApp, IMyService) for the ServerApp and ServerClass. ClientApp is a Windows Form, and it implements IClientApp. ClientClass holds the necessary interfaces (IClientApp, IClientClass) for the ClientApp and ClientClass. LinkedList is a project consisting of a class named LinkedList. I have provided necessary comments in its coding. Please refer to it now.
ServerClass
IServerApp
IMyService
ClientApp
IClientApp
ClientClass
IClientClass
LinkedList
In the Server textbox, enter the machine name on which the server is running. If the client and the server are running on the same machine, you can enter "localhost".
localhost
In the Username textbox, enter the username you want to use. The server checks the username. If the username is in use, it will request you to change the name.
Click Logon. If any of the above textboxes is empty, it will request you to fill the necessary textbox.
You can send the message in two ways:
messageEntryBox
If the "Global" checkbox is not checked, no user will be selected from the user list, now if you try to send a message, a warning will be popped up asking you to select a user.
Well, I haven't made the user list multi-selectable. Multi messaging other than "Global send" is currently not implemented.
The titlebar of the dialog is a custom made titlebar. I have disabled the original Toolbar of the Windows Form. (To do this, enter the properties of the Windows Form, change the value of the ControlBox property to false and also delete the value of the Text property.) I have put a label and a button at the top of the form to act as a titlebar. I have also changed their anchor settings for keeping their size and position synchronized with the changes in the main window. I have added an event for the label to move the window as expected from the original titlebar. (Thanks to MinaFawzi's article Creating a non rectangular form using GDI+.
ControlBox
false
Text
I have also set the anchor for all the controls, so when the main window is resized they arrange themselves in accordance with the new situation.
If you click the button "X", the client logs off the server and will be closed. If you click "Logoff", a message box pops up asking for your choice...to logoff without closing the window, to logoff and close the window and do nothing... A better message box can be arranged here, but I don't have that much time.
For further improvements, an icon can be placed in the system tray and the taskbar display can be disabled. I may implement this later on.
During the early stages of the coding process, I first created a server and made the clients regularly check the server for updates. But it seemed to me a little bit silly. Then I searched for another approach. The correct approach should be "a client sends a message to someone, the server gets the message and forwards it to the concerned user"; but how? So I decided to design the client side like the server side (in fact registering a separate TCP channel for each client and creating the client side class object that can be called by the server for pushing the message to the client). The main design is the same for both the server and the client sides. Create a class object with new and then marshal it.
new
ClientClass remService = new ClientClass();
ObjRef obj = RemotingServices.Marshal(remService,"TcpClient");
ServerClass remService = new ServerClass();
ObjRef obj = RemotingServices.Marshal(remService,"TcpService");
(Well, I must confess that .NET really facilitates communication between remote applications. If we develop this project in C++ using COM, lots of things will be coded manually and longer time will be required.) But later-on, again a change is required. Because the program will run on a network terminal server, the clients will log onto the network server with a terminal client. In the first approach, all the clients have the same static port number. This causes no problem for individual clients running on different machines, but in the terminal server-terminal client approach this static port causes problems... To overcome this situation, I have found a solution: before registering a TCP channel, the client can ask the server to send the port number. After getting this port number, the client registers its TCP channel and on the server side the server keeps this information to find the appropriate clients later on. The server sends the unique TCP number to each registering client. As a result, multiple clients can be initiated on a single and/or multiple machine(s).
Also I'd like to point out a problem that hampered me a lot. When I first designed the client, it always hanged on when I pressed the logon button to logon to the server. I couldn't find a solution at first, but then I found this article from Ingo Rammer: Thinktecture. It helped me a lot.
I have tested this client-server solution in a network environment with multiple machines running on WinXP Professional SP2. (.NET Framework 1.1). Much more testing is required, but this is all for now.
I tested the solution on VS 2005 and Vista. | http://www.codeproject.com/Articles/10599/A-Chat-Server-Client-Solution-for-Local-Networks | CC-MAIN-2014-49 | refinedweb | 1,321 | 64.41 |
Principle of quick sorting:
Select a key value as the base value. Those smaller than the benchmark value are all on the left (generally unordered), and those larger than the benchmark value are all on the right (generally unordered). Generally select the first element of the sequence.
Compare from the back to the front, and use the benchmark value to compare with the last value. If the exchange position is smaller than the benchmark value, if the next value is not compared, the exchange will not be started until the first value smaller than the benchmark value is found. After finding this value, compare from the front to the back. If there is a value larger than the reference value, exchange the position. If you do not continue to compare the next value, do not exchange until you find the first value larger than the reference value. Until compare index from front to back > compare index from back to front, end the first cycle, at this time, for the benchmark value, the left and right sides are orderly.
Then compare the left and right sequences respectively and repeat the above cycle.
Code instance:
public class FastSort { public static void main(String[] args) { System.out.println("Quick sort test"); int[] a = { 121, 16,12,222,3333,212, 15, 1, 30,23, 9,33,56,66,543,65,665 }; int start = 0; int end = a.length - 1; sort(a, start, end); for (int i = 0; i < a.length; i++) { System.out.println(a[i]); } } public static void sort(int[] a, int low, int high) { int start = low; int end = high; int key = a[low]; while (end > start) { // Compare back to front while (end > start && a[end] >= key) // If there is no one smaller than the key value, compare to the next until there is a swap location smaller than the key value
//And then compare before and after end--; if (a[end] <= key) { int temp = a[end]; a[end] = a[start]; a[start] = temp; } // Compare before and after while (end > start && a[start] <= key) // If there is no larger than the key value, compare to the next until there is a swap location larger than the key value start++; if (a[start] >= key) { int temp = a[start]; a[start] = a[end]; a[end] = temp; } // At the end of the first cycle comparison, the position of the key value has been determined.
// The values on the left are smaller than the key values, and the values on the right are larger than the key values
// But the order of the two sides may be different. Make the following recursive call } // recursion if (start > low) sort(a, low, start - 1);// Left sequence. First index location to key index-1 if (end < high) sort(a, end + 1, high);// Right sequence. Index from key+1 To the last } }
| https://programmer.group/java-quick-sort-method.html | CC-MAIN-2020-24 | refinedweb | 472 | 65.35 |
Question:
I was wondering how to achieve the following in python:
for( int i = 0; cond...; i++) if cond... i++; //to skip an run-through
I tried this with no luck.
for i in range(whatever): if cond... : i += 1
Solution:1
Python's for loops are different.
i gets reassigned to the next value every time through the loop.
The following will do what you want, because it is taking the literal version of what C++ is doing:
i = 0 while i < some_value: if cond...: i+=1 ...code... i+=1
Here's why:
in C++, the following code segments are equivalent:
for(..a..; ..b..; ..c..) { ...code... }
and
..a.. while(..b..) { ..code.. ..c.. }
whereas the python for loop looks something like:
for x in ..a..: ..code..
turns into
my_iter = iter(..a..) while (my_iter is not empty): x = my_iter.next() ..code..
Solution:2
There is a
continue keyword which skips the current iteration and advances to the next one (and a
break keyword which skips all loop iterations and exits the loop):
for i in range(10): if i % 2 == 0: # skip even numbers continue print i
Solution:3
Remember that you are iterating over the elements in the list, and not iterating over a number.
For example consider the following:
for i in ["cat", "dog"]: print i
What would happen if you did i+1 there? You can see now why it doesn't skip the next element in the list.
Instead of actually iterating over all values, you could try to adjust what is contained inside the list you are iterating over.
Example:
r = range(10) for i in filter(lambda x: x % 2 == 0, r): print i
You can also consider breaking up the for body into 2. The first part will skip to the next element by using
continue, and the second part will do the action if you did not skip.
Solution:4
You can explicitly increment the iterator.
whatever = iter(whatever) for i in whatever: if cond: whatever.next()
You will need to catch StopIteration if cond can be True on the last element.
Solution:5
There is an alternate approach to this, depending on the task you are trying to accomplish. If
cond is entirely a function of the input data you are looping over, you might try something like the following:
def check_cond(item): if item satisfies cond: return True return False for item in filter(check_cond, list): ...
This is the functional programming way to do this, sort of like LINQ in C# 3.0+. I'm not so certain it's all that pythonic (for a while Guido van Rossum wanted to remove filter, map and reduce from Python 3) but it certainly is elegant and the way I would do it.
Solution:6
You can't trivially "skip the next leg" (you can of course skip this leg with a
continue). If you really insist you can do it with an auxiliary
bool, e.g.
skipping = False for i in whatever: if skipping: skipping = False continue skipping = cond ...
or for generality with an auxiliary
int:
skipping = 0 for i in whatever: if skipping: skipping -= 1 continue if badcond: skipping = 5 # skip 5 legs ...
However, it would be better to encapsulate such complex looping logic in an appropriate generator -- hard to give examples unless you can be a bit more concrete about what you want though (that "pseudo-C" with two presumably 100% different uses of the same boolean
cond is REALLY hard to follow;-).
Solution:7
for i in filter(lambda x:x!=2,range(5)):
Note:If u also have question or solution just comment us below or mail us on toontricks1994@gmail.com
EmoticonEmoticon | http://www.toontricks.com/2019/02/tutorial-python-for-loop-question.html | CC-MAIN-2019-30 | refinedweb | 615 | 61.16 |
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Re: Hash table performance
From:
Tom Anderson <twic@urchin.earth.li>
Newsgroups:
comp.lang.java.programmer
Date:
Sat, 21 Nov 2009 19:44:25 +0000
Message-ID:
<alpine.DEB.1.10.0911211853410.26245@urchin.earth.li>
This message is in MIME format. The first part should be readable text,
while the remaining parts are likely unreadable without MIME-aware tools.
---910079544-686117080-1258832665=:26245
Content-Type: TEXT/PLAIN; charset=ISO-8859-1; format=flowed
Content-Transfer-Encoding: 8BIT
On Sat, 21 Nov 2009, Marcin Rze?nicki wrote:
On 21 Lis, 19:33, Jon Harrop <j...@ffconsultancy.com> wrote:
I'm having trouble getting Java's hash tables to run as fast as .NET's.
Specifically, the following program is 32x slower than the equivalent
on .NET:
? import java.util.Hashtable;
? public class Hashtbl {
? ? public static void main(String args[]){
? ? ? Hashtable hashtable = new Hashtable();
? ? ? for(int i=1; i<=10000000; ++i) {
? ? ? ? double x = i;
? ? ? ? hashtable.put(x, 1.0 / x);
? ? ? }
? ? ? System.out.println("hashtable(100.0) = " + hashtable.get(100.0));
? ? }
? }
My guess is that this is because the JVM is boxing every floating point
number individually in the hash table due to type erasure whereas .NET
creates a specialized data structure specifically for a float->float hash
table with the floats unboxed. Consequently, the JVM is doing enormously
numbers of allocations whereas .NET is not.
Is that correct?
You are using Hashtable instead of HashMap - probably the performance
loss you've observed is due to synchronization (though "fat"
synchronization might be optimized away in case of single thread you
still pay the price, though lower). If you took a look at JavaDoc, you'd
notice that HashTable methods are synchronized As of boxing, you are
correct (though there is no type erasure in your example because you did
not specify type parameters at all) but I suspect that these costs are
not the most contributing factor to overall poor performance. I'd blame
synchronization in the first place.
I'd be *very* surprised if that was true. In this simple program, escape
analysis could eliminate the locking entirely - and current versions of
JDK 1.6 do escape analysis. Even if for some reason it didn't, you'd only
be using a thin lock here, which takes two x86 instructions and one memory
access for each lock and unlock operation, far less than the boxing or
unboxing.
I modified the test code to look like this (yes, with no warmup - this is
very quick and dirty):
import java.util.Map;
import java.util.HashMap;
import java.util.Hashtable;
public class HashPerf {
public static void main(String args[]) throws InterruptedException{
for(int i=1; i<=100; ++i) {
long t0 = System.nanoTime();
test();
long t1 = System.nanoTime();
long dt = t1 - t0;
System.out.println(dt);
System.gc();
Thread.sleep(200);
}
}
private static void test(){
Map<Double, Double> hashtable = new HashMap<Double, Double>();
// Map<Double, Double> hashtable = new Hashtable<Double, Double>();
for(int i=1; i<=1000; ++i) {
double x = i;
// synchronized (hashtable) {
hashtable.put(x, 1.0 / x);
// }
}
}
}
And then ran it with three variations on the comments: one as above, one
uncommenting the synchronization of the hashtable, and one switching the
HashMap to a Hashtable. I have java 1.5.0_19 on an elderly and ailing
PowerPC Mac laptop. I ran with -server and otherwise stock settings.
The timings for each show the usual power curve distribution: 80% of the
measurements are no more than 50% longer than the fastest, and 90% are no
more than twice as long, with the last 10% being up to 10 times longer. If
we say that the slowest 10% are artifacts of warmup, GC, the machine doing
other things, etc, and ignore them, then the average times i got were
(with standard error of the mean, which is broadly like a ~60% confidence
limit IIRC):
HashMap 933500 +/- 15006
sync HashMap 1003200 +/- 16187
Hashtable 868322 +/- 11602
That is, adding synchronization to the accesses adds a 7.5% overhead.
Although somehow, the old Hashtable comes out faster!
So, even with java 1.5, adding synchronization to HashMap.put() imposes
only a small performance penalty - i'd expect it to be less with 1.6. I
doubt very much that this is the major factor in the OP's performance
problem.
tom
--
.... the gripping first chapter, which literally grips you because it's
printed on a large clamp.
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Hi, i'm new to C and i need a little help with a function i am writing to read in user input from the keyboard, a set of integers, and store them into an array. The input is terminated when the user types any invalid integer input. Integers are seperated by white space or new lines. Here is what I got so far... Do I use the ISDIGIT function as a condition for the loop? Also the (empty) array is created in the main program, and passed over to the function to add to it.
--------------------------------------------------------------------
#include <stdio.h>
void get_input(char input_array[]);
int main()
{
char input_array[999];
get_input(input_array);
return 0;
}
void get_input(char the_array[])
{
int i;
for (i=0; ???????????; i++) {
printf("Enter stream of integers: ");
scanf("%c", &the_array[i]);
}
}
--------------------------------------------------------------------
I hope there aren't TOO many mistakes in there
If anyone can give me any advice in how to create the condition for the loop, it would be much appreciated!If anyone can give me any advice in how to create the condition for the loop, it would be much appreciated!
Thankyou! | http://cboard.cprogramming.com/c-programming/14965-isdigit-inputs-numbers-into-array-til-user-types-invalid-integer.html | CC-MAIN-2016-30 | refinedweb | 184 | 69.01 |
Posted in category machine learning with tags definition.
- What’s Machine Learning?
- Differences between supervised learning / unsupervised learning?
- What is overfitting and underfitting?
- What’s the train/test split’s problem?
- What is different between training set / test set / cross validation set?
- The idea of K-Nearest Neighbours (KNN)?
- What’s a decision tree?
- Differences between logistic regression and linear regression and other classification algorithms?
- The idea of Support Vector Machine (SVM)?
- Differences between Clustering and Classification
- K-means Clustering?
This is the first post in the series of posts I wrote about what I’ve learned in Machine Learning and Data Science on the online courses. I write these posts by supposing that there are some interviewers asking me some technical questions in these fields. I also write the answers in the context where someone doesn’t know anything about ML or Data ask about these fields.
This article is not for you to learn, it’s for refrence only!
What’s Machine Learning?
A computer program is said to learn from experience E with respect to some class of tasks T and performance measure P, if its performance at tasks in T, as measured by P, improves with experience E. – Tom Mitchell.
Above definition was mentioned in the course of Andrew NG in his course on Coursera. Shorter, for me, ML is the science in that we give computers abilities to act without being explicitly programmed. They can learn from some given data (they can even collect data by themselves), using some algorithms and make the prediction.
Differences between supervised learning / unsupervised learning?
- Supervised learning : output is already known.
- Regression : linear regression, logistic regression, …
- Classification : K-Nearest Neighbors (KNN), Support Vector Machine (SVM), Decision Tree
- Unsupervised learning: We have very little or no idea what our results should look like (Clustering)
What is overfitting and underfitting?
- Underfitting : Capturing not enough pattern in the data. The curve under fit the data points. The model performs poorly both in the training set and test set.
- Overfitting : Contrary to underfitting, capturing noise, unnecessary patterns which do not generalize well to unseen data. The model performs very well on the training set but poorly on the test set.
Other names : overfitting (high variance), underfitting (high bias).
What’s the train/test split’s problem?
It depends highly on the way we choose the train/test set data. That’s why we need to use cross-validation evaluation to fix it. For example, K-fold cross validation.
# Split arrays or matrices into random train and test subsets from sklearn.model_selection import train_test_split X_train, X_test, y_train, y_test = train_test_split(X, y, test_size=0.33, random_state=42)
If we have enough data, it may be a good choice.
What is different between training set / test set / cross validation set?
- Training set : Your algorithms adjust their parameters based on training data.
- Validation set : Your algorithms are run on validation set to get a result. This result is compared to other training/validation data to choose the best one.
- Cross validation set : We don’t split the data into fixed partitions, hence CV.
- Above things give us the reason why we need to choose CV and training sets on the same distribution so that we can compare between them.
- Test set : You run the “winner” algorithm (parameters) on the test set to see how well your model works in the real world.
The idea of K-Nearest Neighbours (KNN)?
- This is a classification algorithm, we choose a categories for some example based on the category of their neighbors. For example, if $K=1$, the category of that example is the same with the category of the nearest example to it.
- Calculate the accuracy with different numbers of $K$ and then choose the best one.
from sklearn.neighbors import KNeighborsClassifier neigh = KNeighborsClassifier(n_neighbors=3) neigh.fit(X, y) print(neigh.predict([[1.1]])) print(neigh.predict_proba([[0.9]]))
What’s a decision tree?
- The basic intuition behind a decision tree is to map out all possible decision paths in the form of a tree.
- We need to choose the best attribute with the highest significance and split the data based on that attribute. But what is the measure to choose this attribute?
- We base on the entropy and the information gain.
Entropy (of each node, between 0 and 1) : the amount of information disorder or the amount of randomness in the data. The lower the entropy, the less uniform the distribution, the purer the node. Each node (entropy) have several ways to split the data (based on different attributes)
Information gain : The information we “gain” after the split. The tree with the higher information gain will be chosen.
from sklearn.tree import DecisionTreeClassifier drugTree = DecisionTreeClassifier(criterion="entropy", max_depth = 4) drugTree.fit(X_trainset,y_trainset) predTree = drugTree.predict(X_testset)
Differences between logistic regression and linear regression and other classification algorithms?
We use Logistic Rregression instead of Linear Regression because :
- The dependent variable is binary / multiclass (categorical variable), we cannot apply LR in this case.
- We need the know the probabilities of the results.
We use Logistic Regression instead of other categorical classifiers because :
- The predicted probabilities in LR are more well-calibrated.
- If we need to see the impact of some features on the dependent variables.
- When we wanna check the decision boundary.
from sklearn.linear_model import LogisticRegression LR = LogisticRegression(C=0.01, solver='liblinear').fit(X_train,y_train) # C parameter indicates inverse of regularization strength which must be a positive float. yhat = LR.predict(X_test) # get the probability # the first column is the probability of class 1, P(Y=1|X), and second column is probability of class 0, P(Y=0|X): yhat_prob = LR.predict_proba(X_test)
The idea of Support Vector Machine (SVM)?
- It’s a supervised algorithm that classifies cases by mapping data to higher-dimensional feature space and finding the separator.
- How we can transform data to the higher-dimensinal space? By using kernels.
- How we can find the best separator after the transformation? By using SVM.
- Find the largest separation (margin) between two classes.
from sklearn import svm clf = svm.SVC(kernel='rbf') clf.fit(X_train, y_train) yhat = clf.predict(X_test)
Differences between Clustering and Classification
- Classification : you have labeled data and need to make a prediction on the unknown object. It’s supervised learning.
- Clustering : you working on unlabeled data and need to split data into groups where objects have the common properties. It’s unspervised learning.
K-means Clustering?
- We need to separate the data into different groups without knowing their labels beforehand.
- The process is based on “K” centroids at the first glance (they are chosen randomly).
- We divide the data based on the distance of each point to the sigmoids. The shortest distance decides the group to which that point belongs.
- After we have the frist groups, we move the centroids to the mean of each cluster and start again the algorithm until we don’t have too much change on the centroids.
- We need to do several differently initial centroids to find the best results and avoid the local optimization.
- After trying with different K, we need to apply the elbow method to choose the most suitable K.
from sklearn.cluster import KMeans k_means = KMeans(init = "k-means++", n_clusters = 4, n_init = 12) # initial centroids # k-means++ : k-means clustering in a smart way to speed up the convergence # n_init : number of times we "re-choose" the initial centroids. k_means.fit(X) k_means_labels = k_means.labels_ # labels for points k_means_cluster_centers = k_means.cluster_centers_ # coordinates of the cluster centers
Source of figures used in this post: overfitting, decision tree. | https://dinhanhthi.com/what-is-machine-learning | CC-MAIN-2019-22 | refinedweb | 1,263 | 50.23 |
Thursday, June 02, 2016
read again the How to Use Emacs section from Clojure for the Brave and True and cider documentation: Using the REPL, learned a few new tricks:
emacs:
M-\: Delete all spaces and tabs around point.
M-.and
M-,: Navigate to source code for symbol under point and return to your original buffer.
cider
C-enter: Close parentheses and evaluate.
C-c C-o: Remove the output of the previous evaluation from the REPL buffer.
C-c C-u: Kill all text from the prompt to the current point.
C-c C-band
C-c C-c: Interrupt any pending evaluations.
docs
C-c C-d C-d: Display documentation for symbol under point.
C-c C-d C-a: Apropos search for functions/vars.
C-c C-d C-f: Apropos search for documentation.
C-c C-d j: Display JavaDoc.
C-c M-i: Inspect expression.
history
to save repl to a file after
cider-quit
(setq cider-repl-history-file "path/to/file")
to assign last eval to a variable, use
*1,
*2 ...,
*e for last exception
(def x *1)
Friday, June 03, 2016
not just me have slow
cider-jack-in problem: Do some profiling to find out why it takes so long to start a REPL · Issue #1717
porting some java stuffs to kotlin, it's such a pleasant to work with. Emacs doesn't have any kotlin support, I just use intellij idea, not bad after I figured out how gradle works under idea. I like that idea will convert to kotlin codes when I pasted java codes into a kotlin file, it's not 100%, but basic struture is there and much better than write from scratch.
Saturday, June 04, 2016
I love ncurses apps, I'm using
mc and
cmus daily.
today I just did a
apt-cache search ncurses, tried few more apps:
- tig, a git browser, pretty good.
- iptraf-ng, network monitoring tool.
- nast, network analyzer sniffer tool
- cpm: console password manager
- lnav: log file navigator
- iselect: interactive selection tool
- multitail: view multiple log files
- sinfo: monitoring tool for networked computers
- wordgrinder-ncurse: word processor
Sunday, June 05, 2016
found quite good looking font for terminal: F25 Bank Printer Font
more networking books to read, borrowed another classic: Effective TCP/IP Programming: 44 Tips to Improve Your Network Programs from local library.
Monday, June 06, 2016
java convesion between string and timestamp, without using
clj-time:
(import 'java.sql.Timestamp 'java.text.SimpleDateFormat) (defn ts->str [ts format] (->> ts (* 1000) Timestamp. (.format (SimpleDateFormat. format)))) (defn str->ts [str format] (->> str (.parse (SimpleDateFormat. format)) .getTime (* 0.001) long))
Tuesday, June 07, 2016
worked out something to manage dependencies of a single
clj file.
when I want to start something quick or just test something on clojure, simply create a single
clj feels more lightweight than
lein new and
cider. but I can't use external libraries in this way, tools like lein-try and lein-oneoff helped not really ideal for scripting.
I checked how to download dependencies,
mvn dependency:copy-dependencies is the easiest way I found, even though it requires maven ...
I wrapped up the tool and put it on github: usrjim/oneclj
Wednesday, June 08, 2016
the latest cognicast is Clojure spec with Rich Hickey - Cognicast Episode 103, the part about changes and compatible is fascinating.
however, one thing I worried is how much performance tradeoffs
clojure.spec makes? and people are benchmarking it
kotlin code can be quite concise, I changed from
fun filteredBy(region: String, fn: (resultData) -> Boolean): resultData? { val result = getResult(region)?.filter(fn) return result?.first() }
to
fun filteredBy(region: String, fn: (resultData) -> Boolean): resultData? { return getResult(region)?.filter(fn)?.first() }
and finally
fun filteredBy(region: String, fn: (resultData) -> Boolean): resultData? = getResult(region)?.filter(fn)?.first()
to get recent browsed urls from firefox:
cat ~/.mozilla/firefox/xxxxxx.default/sessionstore-backups/recovery.js | jq '.windows[].tabs[].entries[]|.title+","+.url'
sharing files within an internal network, can simply start php server inside the folder:
php -S 0.0.0.0:3000
if using clojure, with my oneclj:
;;[ring/ring-core "1.5.0-RC1"] ;;[ring/ring-jetty-adapter "1.5.0-RC1"] (use 'ring.adapter.jetty) (use 'ring.middleware.file) (-> {} (wrap-file ".") (run-jetty {:port 3000}))
Friday, June 10, 2016
I always want to know what Ruby Rogues' intro music is. Today when I'm listening 3 Doors Down - Kryptonite, I laughed: this is the ruby rogues music.
the clj above was using
ring 1.5.0-RC1, funny that 1.5.0 was released on that day as well
it will be nice that my oneclj could pack the
clj as a
jar. I think all I need to figure out is the manifest file
a simple poc version:
hello.clj
(ns hello (:gen-class)) (defn -main[] (println "fat jar!"))
compile with:
#!/usr/bin/env bash mkdir -p classes lib cp /opt/clojure/clojure-1.8.0.jar lib/ java -cp .:lib/* clojure.main -e "(compile 'hello)" cat <<EOF > manifest.txt Main-Class: hello Class-Path: lib/clojure-1.8.0.jar EOF jar cfvm hello.jar manifest.txt -C classes . lib
run with:
java -jar hello.jar
but seems there're some limitation: the
ns,
:gen-class and
-main function are required. not easy to wrap a
clj file with these by scripting. let's see.
Saturday, June 11, 2016
an updated poc version of packing a
clj as a
jar, it's working and I'm quite happy with it.
#!/usr/bin/env bash jars=$(find lib/ -type f | sed 's/^/ /') clojure=/opt/clojure/clojure-1.8.0.jar mkdir -p classes lib cp $clojure lib/ cat <<EOF > hello.clj (ns hello (:gen-class)) (defn -main[] (load-file "sample.clj")) EOF java -cp .:$clojure clojure.main -e "(compile 'hello)" cat <<EOF > manifest.txt Main-Class: hello Class-Path: $jars EOF jar cfvm hello.jar manifest.txt -C classes . lib sample.clj
will wrap it up and commit to the oneclj repository.
Sunday, June 12, 2016
so sad the way that packing stuffs in a
jar does not work.
it did work because I got my dependencies/script in the same relative location so that the program could reach to those files.
it doesn't work because:
- put jars in manifest
Class-Pathwon't be loaded, unless you have custom code to load all those dependencies.
- unjar all dependencies and pack them into a
jarwon't work, because you'll need to compile all clojure dependencies (AOT).
even use maven or one-jar I wonder whether they could compile clojure libraries. so I will drop it,
lein and
boot are the right tools to handle this problem.
lots of readings about My Increasing Frustration With Clojure and discussions on reddit and proggit
Monday, June 13, 2016
I was thinking about the jar packing problem while taking the train. coz I remembered
clojure-slim.jar is an un-compiled version of
clojure.jar, which means
clojure.jar is already compiled to java classes and should be able to called within the jar.
also once
clojure has been bootstraped, it won't be matter whether or not other clojure libraries are compiled, they can be interpreted by the main
clojure library.
I suspected it didn't work because I included two
clojure.jar:
clojure-1.8.0.jar (by me) and
clojure-1.7.0.jar (by maven).
still have another problem, how to load the
clj file inside a jar?
I was using load-file but surely it can't refer to a file inside the jar.
load-script is the correct function to use. (note, the namespace is
clojure.main not
clojure.core), using
/@sample.clj can call
sample.clj inside a jar successfully.
so decided to try again, removed
clojure-1.7.0.jar, unpacked all the jar dependencies and included the classpath in the manifest, changed
load-file to
clojure.main/load-script.
it worked. the generated
jar file ran everywhere.
the fat-jar bash script was committed to usrjim/oneclj
when googling how to run
clj scirpt, accidentally found this page and it has useful information about Enhancing Clojure REPL with rlwrap
this stackoverflow post has information for cleaning the repl?
to set multiple lines variable in a
Makefile, can do with:
define mline = line one line two some $(VAR) endef
calling with
$(mline) will treat each line as a command, a trick I found is:
export mline @echo "$$mline"
Tuesday, June 14, 2016
added the
rlwrap enhancement to oneclj, will generate completion list for auto-completion.
however, it requires
use the namespace first coz now simply take functions from
(all-ns). I wanted to include all functions in all dependencies, but seems another difficult task, maybe enhance later.
one project I wanted to do is analyze my photos and tag them for search. it's a deep topic, of course I will use 3rd party library of API.
google, ibm, microsoft are all doing it, can find more information in this hacker news post.
but I don't want to send my photos to them for privacy reason, I'm looking for a open-source solution. TensorFlow looks big enough and may worth spending some time with. and their image recognition tutorial does exactly what I need.
maybe need to read the book first?: Neural networks and deep learning
Wednesday, June 15, 2016
try tensorflow later until there is java support (currently python and c++)
for a comparison of Deeplearning4j vs. Torch vs. Theano vs. Caffe vs. TensorFlow, just pick one of them to try.
reading Neural networks and deep learning, having some difficulties already, but still able to follow, see if I could finish the first example.
when working with clojure on one of my old laptop, the executing speed is really slow. i turned on
-verbose and it took a very long time to load all the libraries (on my old and slow laptop). it was just a simple script and won't need most of the libraries to be loaded.
there're tools to shrink library size, like ProGuard, but again, make them work with clojure gonna be very difficult.
there're also tools to reduce jvm startup time, drip, nailgun, etc.
drip is easy to use and uses fresh jvm, it enhanced about 50% startup time when I tested.
however, even re-use jvm, calling
clojure.main to evaluate a script still quite slow, I guess there will be something to do for improvement, from the page I saw AOT all libraries actually just helped a little bit of the speed even though AOT is suggested, not worth the time to try making the script AOT for now.
found a useful package for emacs: clj-refactor.el
the demo gif is adding a core library
clojure.string, I wonder how it works on other dependencies.
well, there is a new build/deployment tool: Habitat, by Chef
got attentions because it's written in rust
this is my frustration with all these containerization things. I'm frustrated because container is a good thing and we should adapt it. but new tool/framwork every month? (talking about those who did better work than existing ones).
I would rather spend time for LXC and read some docker source code.
anyways, I'm reading Puppet for Containerization, quite good, learned few new things.
first, you can define multiple vagrant machine in a yaml config file: see Multi-Machine Vagrant with YAML
it's a very simple way to create multi-nodes local environment.
then it's couple protocols, RAFT and Gossip
and the docker VXLAN-based overlay network and more details (require linux kernel
>=3.19)
wait, what is VXLAN?
note, upgrade kernel from ubuntu trusty (
3.13.0), install one of these packages:
linux-generic-lts-vivid(
3.19.0)
linux-generic-lts-wily(
4.2.0)
linux-generic-lts-xenial(
4.4.0)
using the latest
16.04 LTS is easier.
the stack of examples in the book also includes registrator for service registry, BIND for dns, consul for service discovery and key-value store, swarm for docker clustering, and of course, puppet for configuration management.
going through the examples is easy, but when applying to your own stack is a different story. it involves so many tools and digging deep into each of those area is a huge task.
the stask looks reasonable and this book helps a lot for getting a general idea of containerization. actually I'm more interested in this kind of DIY appoaches than kubernetes.
Thursday, June 16, 2016
Lamport timestamps is the lamport clock thing mentions in gossip protocol, for distributed message ordering.
the protocol also mentions the payload and entire message framing must fit within a single UDP packet
so I googled how large is a single UDP packet: The most reliable and efficient udp packet size?
it's the minimum MTU size that an host can set is 576 and IP header max size can be 60 bytes (508 = 576 MTU - 60 IP - 8 UDP)
576 bytes is the RFC791, MTU minimum packet size
1500 bytes is the MTU of Ethernet limit (some mathematics of Ethernet here)
so fit within a single UDP packet to avoid fragments: UDP Packet size and packet losses
some random reads:
a discussion on Why isn't there an open sourced Datomic?, brings out hitchhiker-tree: a datomic like database.
ClickHouse: a distributed column-oriented DBMS
Ancient code: programming on paper!
My First 10 Minutes On a Server - Primer for Securing Ubuntu, usually if this kind or article gets on frontpage, the comments are more interesting. discussion on hacker new and proggit.
learned couple things from the comments: Port knocking and Should I change the SSH port to < 1024?
Friday, June 17, 2016
did a super simple euro 2016 fixtures and group table page using this api
a simple function to convert UTC to lcoal timezone (
+8)
(require '[clj-time.core :as t]) (require '[clj-time.format :as f]) (defn to-local-datetime [d df] (let [parsed-date (f/parse d) my-timezone (t/time-zone-for-offset 8) date-format (f/formatter-local df)] (f/unparse date-format (t/to-time-zone parsed-date my-timezone)))) (to-local-datetime "2016-06-17T21:03:57Z" "EEE MMM dd hh:mm") ;; "Sat Jun 18 05:03"
the free-keys emacs package lists out all unused keybindings (under current mode) for you.
I decided change my terminal escape key to
M-[, since
M-n is for auto-complete selection.
Saturday, June 18, 2016
quickly went through the rest of Puppet for Containerization
later chapters are examples of how to implement a specific setup, so lots of configuration codes and not much new things could be found. one example uses kubernetes, but a very brief introduction only.
the take away is there is docker and kubernetes modules for puppet.
continue reading Java Network Programming, since kotlin also comes with a repl (
./bin/kotlinc-jvm), so it's quite useful to test out examples in the book (convertion from java codes is more straight forward than using clojure):
$ ./bin/kotlinc-jvm Welcome to Kotlin version 1.0.2 (JRE 1.8.0_77-b03) Type :help for help, :quit for quit >>> import java.net.* >>> InetAddress.getByName("") >>> InetAddress.getAllByName("") [Ljava.net.InetAddress;@11be3c5 >>> InetAddress.getAllByName("").forEach(::println) [kotlin.Unit, kotlin.Unit] >>>
one interesting thing is
InetAddress.isReachable():
>>> InetAddress.getByName("").isReachable(1000) false
from Problem with isReachable in InetAddress class, found that you actually need root privilege to run
ping.
learned some history about zip and gzip: How are zlib, gzip and Zip related? What do they have in common and how are they different? - Stack Overflow
working with varnish you'll need to handle the Vary: Accept-Encoding header: you don't want different order of deflate and gzip values create another cache object variant. now I know they are wrappers of zlib, the library browser uses.
and I understand now why most of time you'll see smaller
.tar.gz file than
.zip: gzip applies deflate across all files while zip applies per file.
Tuesday, June 21, 2016
Dockercon 2016 is on, first news is Docker 1.12: Now with Built-in Orchestration! and their Docker Stacks and Distributed Application Bundles (DAB)
so it bundled load balancer, service discovery and key-value store, inner node communication using gRPC and http/2, IPVS for load balancing.
Raft consensus group from this paper: In Search of an Understandable Consensus Algorithm (pdf)
I don't see
v1.12 on their xenial repository yet, but really want to try it, meanwhile reading some articles first:
- More Microservices Bliss with Docker 1.12 and Swarm only
- Docker Engine 1.12 comes with built-in Distribution
trying other text-mode browsers, I just couldn't remember so many keys with
w3m, I want to use arrow keys and menus only.
first I tried Links2.
links2 -g starts a graphic mode browser, I like that. but lack of tab bothers me. so I tried another variant: Elinks, and switched to it completely.
comparing to w3m, elinks is better at:
- has menus (
esc)
- an
os shell, open a shell without exiting elinks
- google search by just entering
g textin location box, it's uri rewriting, very powerful. I learned from here
- super fast navigation by numbering links (
.to toggle)
- support colors (
%to switch)
- keybindings for tabs make more sense:
tto open tab,
cto close,
>/
<for switching
- marks:
mto mark,
'to call
one feature I missed is how easy that w3m sends current url to firefox. but elinks can do it easily too:
add these two lines to
~/.elinks/elinks.conf
set document.uri_passing.firefox = "firefox %c 2>/dev/null" bind "main" "F10" = "frame-external-command"
sometimes I'll use emacs terminal mode, didn't work very well because so many keybindings are not working as expected under terminal mode, for example the
esc key.
tonight I just found out that arrow keys and
del key bring my evil-mode
emacs-state to
nomral-state, I don't know why but somehow useful to me.
Wednesday, June 22, 2016
another example to show how concise kotlin is: a daytime client, here is the java code (quite similar to the code in java network programming)
I tried rewriting it to kotlin:
import java.net.Socket val stream = Socket("time-a.nist.gov", 13).getInputStream() stream.reader().buffered().use { it.lines().forEach(::println) }
on os x, need to compile elinks
v0.12pre6 manually.
v0.11 doesn't work, neither from
brew nor compile.
another fast clojure repl solution, I combined it with
drip, only needs
160ms to start a repl, no dependencies, but still cool.
Thursday, June 23, 2016
Eclipse Neon is released. I like the neon theme (good use of css
box-shadow).
repl is nothing new when you working with clojure, even kotlin and java 9 have it. but this article: How I built my JAVA REPL is interesting to me. (the repo is here )
there is no dependency required, amazing.
JHipster generates spring boot and angularjs project, heard it from this video: Making Java Developers Hip Again
wanna find unicode characters that looks normal english letters? this is a list of confusable characters
after Elixir v1.3 released, Erlang/OTP 19.0 has been released as well. I really want to spend some time on erlang/elixir.
here's an interesting talk I found from the erlang/otp release thread: mnesia + leveldb: liberating mnesia from the limitations of DETS
LevelDB is not new, but I saw couple times from news recently, I decided to check it out:
this is the official documentation, there're two java wrappers: LevelDB JNI and LevelDB in Java
a hello world example (using LevelDB JNI library):
import org.iq80.leveldb.*; import static org.fusesource.leveldbjni.JniDBFactory.*; import java.io.*; class LevelDB { public static void main(String[] args) throws IOException { Options options = new Options(); options.createIfMissing(true); DB db = factory.open(new File("/home/ubuntu/example"), options); try { db.put(bytes("Tampa"), bytes("rocks")); String value = asString(db.get(bytes("Tampa"))); System.out.println("got value: " + value); db.delete(bytes("Tampa")); } finally { db.close(); } } }
and
Makefile:
compile: javac -cp leveldbjni-all-1.8.jar LevelDB.java run: @java -cp .:leveldbjni-all-1.8.jar LevelDB clean: rm -f LevelDB.class
the
leveldbjni-all-1.8.jar is just about 1MB, like a nosql version of sqlite
a full implementation build on top of leveldb is RocksDB by facebook ( more details )
when using
ubuntu/xenial vagrant box, there is some problem with setting the private network. since
15.10 ubuntu changed tranditional network interface names (
eth0,
eth1 ..) to Predictable Network Interface Names
newer version of vagrant resolved this, a workaround is change the
config.vm.network line to
config.vm.network "private_network", ip: "192.168.33.10", auto_config: false config.vm.provision 'shell', inline: "ifconfig enp0s8 192.168.33.10 netmask 255.255.255.0"
ip link gives you all network interfaces in your machine.
npm semantic version calculator is a easy way to get correct version syntax for node.js package.
when
bash-completion is activated, the filename completion does not work with some commands. there is another shortcut to force bash completes filename:
M-/, more details here
Friday, June 24, 2016
Glowroot is an open source application performance management (APM) tool, easy to use. check the demo site
dropwizard metrics is another option, but need other tools for collecting data and dashboard.
rocksdb has another interesting usage - as mysql storage engine, this is the FOSDEM 16 talk: RocksDB Storage Engine for MySQL.
I think leveldb is good for persistent cache and persistent queue, quite interested in working on a simple version of them.
phoenix 1.2 has been released, I've decided I want to try it and will spend some time on it.
adding a memcached cache for the site, why not redis? I don't know.
memcached client for clojure, Spyglass seems the only choice.
it supports binary protocol (some brief introduction I found on google book )
adding it is quite easy, the more difficult problem always be cache invalidation.
I set a long ttl, I know what keys I need to purge when there are updates.
for deployment, I added a simple check to see whether remote git branch has new updates:
git remote update if [ -n "$(git log HEAD..origin/master)" ] then // deploy // purge keys // rebuild cache for those purged keys fi
update value by cache key is not an option, could be done when I added a simple lambda service for storing functions.
an easier way is clear all on deployment:
echo "flush_all" | nc memcache-host 11211 and rebuild some critical cache after purging.
Saturday, June 25, 2016
I'll dive into erlnag/elixir and spend rest of the year on it, as usual, starting with books:
first one is Programming Erlang by Joe Armstrong, he has a blog post about the book. my local library has first edition, I may take a look first.
second one is Learn You Some Erlang for Great Good!, LYAHFGG is a good book, so I think this one probably good as well.
local library also has another one: Introducing Erlang, I think I'll take this one too.
Sunday, June 26, 2016
did a SSL Server Test today and got an F
found that there is Yet Another Padding Oracle in OpenSSL CBC Ciphersuites
an
apt-get update && apt-get upgrade fixed it.
actually I was checking nginx ssl config. I noticed server response time is slower via https comparing to access directly to backend. I didn't expect ssl termination costs so much time, I tried to see is there anything could help from tuning nginx config, found nothing yet.
but I had a long
ssl_ciphers list, so did a little bit tidy;
start reading Learn You Some Erlang for Great Good!, notes will be uploaded here
Monday, June 27, 2016
I'm looking for some cache service implemented by java, found JCS and Ehcache, Ehcache also on the list of awesome java, under distributed applications section, I may check it out first.
also interested in another library under the section: hazelcast, an in-memory data grid.
clojure also has an awesome list, I'm checking with liberator, a library for building restful api. their decision graph (svg) is quite useful even for implementing an api library yourself.
a simple testing app is like:
;;[liberator "0.14.1"] ;;[ring/ring-core "1.5.0"] ;;[ring/ring-jetty-adapter "1.5.0"] ;;[compojure "1.5.1"] (ns libera (:require [liberator.core :refer [resource defresource]] [ring.middleware.params :refer [wrap-params]] [ring.adapter.jetty :refer [run-jetty]] [compojure.core :refer [defroutes ANY]])) (defresource parameter [txt] :available-media-types ["text/plain"] :handle-ok (fn [_] (format "The text is %s" txt))) (defroutes app (ANY "/foo" [] (resource :available-media-types ["text/html"] :handle-ok (fn [ctx] (format "<html>It's %d milliseconds since the beginning of the epoch." (System/currentTimeMillis))))) (ANY "/bar/:txt" [txt] (parameter txt)) (ANY "/" [] (resource))) (def handler (-> app wrap-params)) (run-jetty handler {:port 3000})
Thursday, June 30, 2016
some ssl cert verify commands:
# check openssl x509 -in cert.crt -text -noout # verify openssl x509 -noout -modulus -in cert.crt | openssl md5 openssl rsa -noout -modulus -in private.key | openssl md5 # convert to pem openssl x509 -inform PEM -in cert.crt > cert.pem openssl rsa -in private.key -text > private.pem
using groff to write resume.
-t to preprocess with tbl
-fH to use helvetica fonts (
-fT is times family)
-Tascii outputs ascii text
inside the document, use
.fam H to switch fonts to helvetica
grog helps you determine which preprocessors/macros are required. useful when you copy-paste an example from google search result but don't know which preprocessors should use.
$ grog sample.groff groff -t -ms sample.groff
my last day of this job, it's been 6 years and time for a change.
one good practice I've learned when doing infrastructure is try to make things disposable, so I disposed myself as well.
Blog Archives
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- | https://jchk.net/blog/2016-06 | CC-MAIN-2019-13 | refinedweb | 4,354 | 65.42 |
On Dec 10, 2010, at 01:42 PM, Brian Sutherland wrote: >On Fri, Dec 10, 2010 at 07:17:16AM -0500, Barry Warsaw wrote: >>. > >I originally had it that way, but was strongly advised to change it to >the current method to be able to pass the Debian FTP Master gauntlet. Just goes to illustrate the myth of TOOWTDI. :) AFAICT, from my discussions with various folks on debian-python, it should now[*] be done with the separate binary package. [*] At least until dh_python2, if we can ensure it always DTRT. >The current method using dpkg-divert is not too bad, more packages than >python-zope.interface could include that file as well. So you don't >force installation of python-zope.interface. Doing the diversion means it's harder for someone to explicitly determine which package owns the file. Better (I think!) to have none of them own the file. >It also uses standard dpkg functionality, which is a robustness bonus. True. >> Second, we're going to make a big push after Squeeze is released to convert >> packaging to use dh_python2, the new goodness in Debian Python packaging. > >I've had a brief look at it already and will have a much deeper look >once squeeze is released. AFAIKR the only feature it was missing to >completely cover my usecase was good handling of setuptools extras. Can you be more specific? Before I got swamped with the Python 2.7 transition (it will be the default in Ubuntu 11.04), I began looking at dh_python2, adding unit tests, etc. I do plan to get back to it once we have the archive more happily on Python 2.7[*]. [*] >I'm hoping to be able to completely replace the custom tools we use in >python-zope.* packages with it at some point. +1. We hope to get rid of python-support and python-central. >>. > >Any idea as to the mechanism? Hints, but nothing definite. I'm not sure which part of the tool chain is suppressing the neamespace package's __init__.py, and which part is laying it down when the package gets installed. It *is* nice that the packager doesn't have to worry about it though. I don't think you should have to do the tricks zope.interface is doing, or even do the extra binary package. IOW, it should Just Work. >What about 2nd level namespace packages (horror: zope.app.foo)? Haven't tried that yet. >> Sadly PEP 382 was not complete in time for Python 3.2. > >Yeah, there are a lot of packaging related PEPs coming out lately. It's >really great to see attention being paid to these dusty corners:) Indeed! I think we all owe Tarek and the other folks an unlimited supply of beers at the next Pycon. :) -Barry -------------- next part -------------- A non-text attachment was scrubbed... Name: signature.asc Type: application/pgp-signature Size: 836 bytes Desc: not available URL: <> | https://mail.python.org/pipermail/distutils-sig/2010-December/017157.html | CC-MAIN-2016-50 | refinedweb | 489 | 76.62 |
There was a post about this before:
but no code in it, this is the code I have written and it calculates the sine of x, but not exactly what I was asked to do, I have set counter to 15, and there should be no counter, it should stop when fsum/sum2 < 1e-8. I think this should be done with a do while function, I have tried putting the whole thing inside one but it doesnt work that way, since it finishes the counter first and then goes to the while with the final result. Someone please help.
#include <iostream.h> #include <conio.h> #include <stdio.h> int main() { clrscr(); double x,sum,mult=0,count,sign=1,fsum,count2,mult2=3,sum2=2,fres=0; cin>>x; sum=x; for (count=1;count<15;count=count+2) //first loop for x-x^3+x^5-x^7... { sign=-sign; mult=x*sum; sum=x*mult; fsum=sum; fsum=fsum*sign;//changes sign, 1st one + 2nd one - 3rd one + .... for (count2=count2;count2<count;count2++)//second loop for 3! 5! 7!... { sum2=sum2*mult2; mult2++; } cout<<fsum/sum2<<endl;//-x^3/3! +x^5/5! ... fres=fres+(fsum/sum2); //final result here is sine of x, still gotta add x^1 } cout<<x+fres<<endl;//its +x because at the and gotta add x^1, loop calculates starting //from -x^3 getch(); return 0; } | https://www.daniweb.com/programming/software-development/threads/155511/calculate-sin-x | CC-MAIN-2017-26 | refinedweb | 236 | 71.24 |
Hello friends, hope you all are fine and having fun with your lives. Today’s post is about interfacing of RFID module RC522 with Arduino. RC522 is very simple yet effective module. It is an RFID module and is used for scanning RFID cards. Its a new technology and is expanding day by day. Now-a-days it is extensively used in offices where employees are issued an RFID card and their attendance is marked when they touch their card to rfid reader. We have seen it in many movies that when someone places ones card over some machine then door opens or closes. In short, its a new emerging technology which is quite useful.
I recently get a chance to work on a project in which I have to use RFID reader to scan cards. In this project I have used it for for student attendance so I thought to share it on our blog so that other engineers could get benefit out it.
Let’s first have a little introduction of RFID and then we will look into how to interface RC522 with Arduino. RFID is the abbreviation of Radio frequency identification. RFID modules use electromagnetic fields to transfer data between card and the reader. Different tags are attached to objects and when we place that object in front of the reader, the reader reads that tags.Another benefit of RFID is that it doesn’t require to be in a line of sight to get detected. As in barcode, the reader has to be in the line of sight to the tag and then it can scan but in RFID there’s no such restriction. So, let’s get started with Interfacing of RFID RC522 with Arduino.
You should also read:
Interfacing of RFID RC522 with Arduino.
Now let’s start with the interfacing of RFID RC522 with Arduino. There are many different RFID modules available in the market. The RFID module, which I am gonna use in this project, is RFID-RC522. Its quite easy to interface and works pretty fine. This module has total 8 pins as shown in the below figure:
- SDA
- SCK
- MOSI
- MISO
- IRQ
- GND
- RST
- 3.3V
It normally works on SPI protocol, when interfaced with Arduino board. Interfacing of Arduino and RC522 module is shown in below figure:
- The pin configuration is as follows:
- Now, I suppose that you have connected your RFID module with Arduino as shown in above figure and table, which is quite simple. You just need to connect total 7 pins, IRQ is not connected in our case.
- Now next step is the coding, so first of all, download this Arduino library for RFID RC522 module.
Note:
- Its a third party library, we haven’t designed it, we are just sharing it for the engineers.
Download Arduino Library for RFID
- Now coming to the final step. Upload the below code into your Arduino UNO.
#include <SPI.h>
#include <MFRC522.h>
#define RST_PIN 9
#define SS_PIN 10
MFRC522 mfrc522(SS_PIN, RST_PIN);
void setup()
{
SPI.begin();
mfrc522.PCD_Init();
}
void loop() {
RfidScan();
}
void dump_byte_array(byte *buffer, byte bufferSize) {
for (byte i = 0; i < bufferSize; i++) {
Serial.print(buffer[i] < 0x10 ? ” 0″ : ” “);
Serial.print(buffer[i], HEX);
}
}
void RfidScan()
{
if ( ! mfrc522.PICC_IsNewCardPresent())
return;
if ( ! mfrc522.PICC_ReadCardSerial())
return;
dump_byte_array(mfrc522.uid.uidByte, mfrc522.uid.size);
}
- Now using this code, you can read the RFID no of your card quite easily. Now the main task is to use that number and distinguish them so for that I changed the dump_byte_array function a little, which is given below:
#include <SPI.h>
#include <MFRC522.h>
#define RST_PIN 9
#define SS_PIN 10
MFRC522 mfrc522(SS_PIN, RST_PIN);
int RfidNo = 0;
void setup()
{
SPI.begin();
mfrc522.PCD_Init();
}
void loop() {
RfidScan();
}
void dump_byte_array(byte *buffer, byte bufferSize)
{
Serial.print(“~”);
if(buffer[0] == 160){RfidNo = 1;Serial.print(RfidNo);}
if(buffer[0] == 176){RfidNo = 2;Serial.print(RfidNo);}
if(buffer[0] == 208){RfidNo = 3;Serial.print(RfidNo);}
if(buffer[0] == 224){RfidNo = 4;Serial.print(RfidNo);}
if(buffer[0] == 240){RfidNo = 5;Serial.print(RfidNo);}
Serial.print(“!”);
while(1){getFingerprintIDez();}
}
void RfidScan()
{
if ( ! mfrc522.PICC_IsNewCardPresent())
return;
if ( ! mfrc522.PICC_ReadCardSerial())
return;
dump_byte_array(mfrc522.uid.uidByte, mfrc522.uid.size);
}
- Now using the first code I get the card number for all RFID cards and then in second code I used these numbers and place the check now when first card will be placed it will show 1 on the serial port and so on for other cards.
- So, what you need to do is to use the first code and get your card number and then place them in second code and finally distinguish your cards.
- Quite simple and easy to work.
- Hope I have explained it properly but still if you get any problem ask me in comments.
41 Comments
Hi, I keep getting the errors Arduino: 1.6.6 (Windows 10), Board: “Arduino/Genuino Uno”
RFID_01:22: error: stray ‘\342’ in program
Serial.print(buffer[i] < 0x10 ? � 0" : � “);
^
RFID_01:22: error: stray '\200' in program
RFID_01:22: error: stray '\235' in program
RFID_01:22: error: missing terminating " character
C:\Users\skiidoggy\Documents\Arduino\RFID_01\RFID_01.ino: In function 'void loop()':
RFID_01:16: error: 'RfidScan' was not declared in this scope
RfidScan();
^
C:\Users\skiidoggy\Documents\Arduino\RFID_01\RFID_01.ino: In function 'void dump_byte_array(byte*, byte)':
RFID_01:23: error: expected ':' before 'Serial'
Serial.print(buffer[i], HEX);
^
RFID_01:23: error: expected ')' before ';' token
Serial.print(buffer[i], HEX);
^
exit status 1
stray '\342' in program
i have the same problem. it is showing error “stray ‘342’ in program”.
I had the same problem and found that there is an issue with the font or something during the copy/paste process in regards to the double quotes. So, just remove the 4 double quotes and type them again in your code which should resolve that problem!
This is caused by using Cut and Paste on the code, to following characters are incorrectly copied:
Serial.print(“!”);
should be
Seria.print(“!”);
Note the ” ” the ones copied are unicode characters and cause the error ‘\342’
Is it possible to connect multiple readers to a single arduino? If so how can that be done. I can get one to work just find, once I introduce another I get no response.
Hi,
Yes you canconnect multiple RFID with Arduino. In this case you need to add a virtual SPI port. Add me on Skype and we will discuss it in detail. My skype id is theeenggprojects.
Thanks.
I can not download , please link active
thanks very helpful
I have the same problem. It just takes you to the same website.
Hi! First of all, thanks for the project! I’m trying to start ‘playing’ with Arduino and this RFID module was the chosen one to be the first. I’ve uploaded the code to my UNO but I’m not able to see the card numbers anywhere. Tried to use the Serial Monitor, but nothing happens. How/Where can I see this? Thanks in advance!
You should use the first code given in the article. This code will give you the RFID value of card in your Serial Monitor. In the second code, place that card value in the code and it will place the check.
where exactly in the 2nd code do i put the card no received from the first code?
HI SIR…. I CONT DOWNLOAD THE RFID LIBRARY FILE IN PROTEUS ….PLZ HELP ME
did you solved or not yet ?! because i have the same here
can anyone suggest pin configuration for arduino nano for rfid 522?
in setup I added
Serial.begin(9600); // vince added
Otherwise I got nothing out.
Also made additional changes to help readablity
void dump_byte_array(byte *buffer, byte bufferSize) {
for (byte i = 0; i < bufferSize; i++) {
// Serial.print(buffer[i] < 0x10 ? ” 0″ : ” “);
Serial.print(buffer[i] println
}
Serial.println(); // vince added
}
void RfidScan()
{
if ( ! mfrc522.PICC_IsNewCardPresent())
return;
if ( ! mfrc522.PICC_ReadCardSerial())
return;
Serial.println(“MFRC55 buffer”); // vince added
dump_byte_array(mfrc522.uid.uidByte, mfrc522.uid.size);
}
sir is there any way of scanning of multiple rfid tags with rc522 and arduino uno?
if yes plz tell me the way and code….plz
hello mister, my name is Sony, i cant download header for this program, can you help me?
is there any possibility of scanniing 2 tags simultaneously with rc522?i want to scan multiple tags at a time……
I never tried but i think yeah you can do it, it takes usec to scan so it will be like one after another but real quick.
I got error:’getFingerprintIDez’ was not declared in this scope
while(1){getFingerprintIDez();}
I am new in this AdrIUno
Cheers
Why doesn’t the download button work? If somebody gave the link it would be helpful.
The download link is down, please link active. Thank’s!
where is the library , only page reload when i click on the download button
I need wifi 60 relay controll circuit
with cod
can you help me ????????
Yeah add me on Skype and we will discuss it out.
In your second code can please tell me that where to place the id of my tag ?
Hi, i have this error message; can you help me?, please
Arduino:1.8.3 (Windows 10), Tarjeta:”Arduino/Genuino Uno”
C:\Users\Juan Figueroa!\Documents\Arduino\LECTOR\lector2\lector2.ino: In function ‘void dump_byte_array(byte*, byte)’:
lector2:29: error: ‘getFingerprintIDez’ was not declared in this scope
while(1){getFingerprintIDez();}
I got error:’getFingerprintIDez’ was not declared in this scope
while(1){getFingerprintIDez();}
unable to find you on skype
exit status 1
‘getFingerprintIDez’ was not declared in this scope
what to do for this
Arduino: 1.8.4 (Windows 10), Board: “Arduino/Genuino Uno”
C:\Users\admin\Documents\Arduino\sketch_aug31a\sketch_aug31a.ino:2:21: fatal error: MFRC522.h: No such file or directory
#include
^
compilation terminated.
exit status 1
Error compiling for board Arduino/Genuino Uno.
This report would have more information with
“Show verbose output during compilation”
option enabled in File -> Preferences.
this error is coming while verifying this code, can u help me with this
Please help me to get the link for downloading RFID reader module library for proteus isis. I’m waiting your response. Thank you.
Please help me to get the link for downloading RFID reader module library for proteus isis. I’m waiting your response.
Hi sir I tried to interface it with gsm sim 900a
,but I am unable to send two message to two different number it’s just send to one number in case of first tag being read,so how can I send in case of each tag being read send SMS to specific number
umm there is no download content
i click the button and it duplicates this site
how to write a no. on the card? ((If its the second code i am unable to figure out hoe you wrote so on the card)) Please help me out!
Hai, i am new to this field. I want to ask how to do this project on proteus because the rc522 component was absent in the proteus lib. Anyone know where to download the rc522 for proteus?
Hi,
It’s quite difficult to design RC522 in Proteus so you have to work on hardware directly. 😛
Thanks.
I am a student and I am looking for the rfid library rc522 for proteus
Hi,
We haven’t designed it yet.
Thanks.
DUINOTECH RFID-RC522
+
ARDUINO UNO v3
PINOUT:
RC522 UNO
VCC 5v
RST pin ~5 (digitial PWM~)
GND GND
MISO pin 12
MOSI pin 11
SCK pin 13
NSS pin 10
Install Library MRFC522
have used most of the included example sketches successfully.
I accidentally used 5v instead of 3.3v due to bad print on the UNO board
and found that it didnt work on 3.3v after HOWEVER: I had tried two seperate
USB lilypad ARDUINO’s and a NANO at 3.3v and all to no avail. 5v was the first time
I saw it work and only time.
Pan0ptiK:DCN3T
hi! is it possible to use the rfid as an input device then the ouput will be sent through sms by a gsm module?if yes, how will the code be?thanks! | https://www.theengineeringprojects.com/2015/08/interfacing-rfid-rc522-arduino.html | CC-MAIN-2019-43 | refinedweb | 2,045 | 74.39 |
#include <wx/app.h>
The wxApp class represents the application itself when
wxUSE_GUI=1.
In addition to the features provided by wxAppConsole it keeps track of the top window (see SetTopWindow()) and adds support for video modes (see SetVideoMode()).
In general, application-wide settings for GUI-only apps are accessible from wxApp (or from wxSystemSettings or wxSystemOptions classes).
Event macros for events emitted by this class:
wxEVT_ACTIVATE_APPevent. See wxActivateEvent.
wxEVT_IDLEevent. See wxIdleEvent.
Get display mode that is used use.
This is only used in framebuffer wxWidgets ports such as wxDFB.
Returns true if the application will exit when the top-level frame is deleted.
Return the layout direction for the current locale or
wxLayout_Default if it's unknown.
Returns a pointer to the top window.
Returns true if the application will use the best visual on systems that support different visuals, false otherwise.
Returns true if the application is active, i.e. if one of its windows is currently in the foreground.
If this function returns false and you need to attract users attention to the application, you may use wxTopLevelWindow::RequestUserAttention to do it.
Called in response of an "open-application" Apple event.
Override this to create a new document in your app.
Called in response of an "open-document" Apple event.
Called in response of an openFiles message.
You need to override this method in order to open one or more document files after the user double clicked on it or if the files and/or folders were dropped on either the application in the dock or the application icon in Finder.
By default this method calls MacOpenFile for each file/folder.
Called in response of a "get-url" Apple event.
Called in response of a "print-document" Apple event.
Called in response of a "reopen-application" Apple event.
May be overridden to indicate that the application is not a foreground GUI application under OS X.
This method is called during the application startup and returns true by default. In this case, wxWidgets ensures that the application is ran as a foreground, GUI application so that the user can interact with it normally, even if it is not bundled. If this is undesired, i.e. if the application doesn't need to be brought to the foreground, this method can be overridden to return false.
Notice that overriding it doesn't make any difference for the bundled applications which are always foreground unless
LSBackgroundOnly key is specified in the
Info.plist file.:
This function is similar to wxYield(), except that it disables the user input to all program windows before calling wxAppConsole::Yield and re-enables it again afterwards.
If win is not NULL, this window will remain enabled, allowing the implementation of some limited user interaction. Returns the result of the call to wxAppConsole::Yield.
Works like SafeYield() with onlyIfNeeded == true except that it allows the caller to specify a mask of events to be processed.
See wxAppConsole::YieldFor for more info.
Set display mode to use.
This is only used in framebuffer wxWidgets ports such as wxDFB.
Allows the programmer to specify whether the application will exit when the top-level frame is deleted.
Allows runtime switching of the UI environment theme.
Currently implemented for wxGTK2-only. Return true if theme was successfully changed.
Sets the 'top' window.
You can call this from within (or better, any wxTopLevelWindow) in its top-level window list, when it needs to use the top window. If you previously called SetTopWindow() and now you need to restore this automatic behaviour you can call
. wxApp instance and won't have any effect when called later on. This function currently only has effect under GTK. | https://docs.wxwidgets.org/trunk/classwx_app.html | CC-MAIN-2019-47 | refinedweb | 613 | 57.98 |
Starting Your First App
This guide walks you through setting up your app with Kinvey.
Prerequisites
- Xcode 8.1 or above.
- iOS 9 or above.
- Swift 3 or above. Refer to the download page for the Kinvey version that matches your preferred Swift version.
Set Up Kinvey
You can start using the Kinvey framework in one of three ways - using CocoaPods, using the Kinvey starter app, or using the SDK source code.
Using CocoaPods
If you are using CocoaPods, add the Kinvey Pod to your target in the Podfile.
pod `Kinvey`
For a target called
MyProject, your Podfile will look like this:
target 'MyProject' do pod 'Kinvey' end
From the Terminal, run
pod install in the project folder to install the dependency.
3.2.x.
pod 'Kinvey', '~> 3.2.1'
Using the Starter App
You can get started with the Kinvey iOS starter app. The starter app is bundled with the Kinvey framework.
We recommend updating the Kinvey framework in the app to the latest available on our downloads page before you begin.
Using SDK Source Code
The Kinvey iOS SDK is open source. If you prefer to compile your app against the SDK source code, you can follow the instructions on our github repo to set up the SDK.
Add an App Backend
In the Kinvey console, click Create an App and enter the name of your app when prompted.
You can find your key and secret in the dropdown menu in the environment sidebar.
Copy the key and secret when performing the next steps.
Initialize Kinvey Client
Before you can communicate with the backend, you need to give the library your app key and secret. This is done with the following code snippet, substituting your app's key and secret (found in the management console header when viewing your app backend). We recommend adding this to your AppDelegate's
application:didFinishLaunchingWithOptions: method.
import Kinvey Kinvey.sharedClient.initialize( appKey: "<#Your App Key#>", appSecret: "<#Your App Secret#>" ) { switch $0 { case .success(let user): if let user = user { print("\(user)") } case .failure(let error): print("\(error)") } } // Then access data store collections on the shared instance let dataStore = try DataStore<Book>.collection()
The above snippet uses the
sharedClient global variable. This shared instance of the
Kinvey.Client class adds convenience when your app is only communicating with a single Kinvey instance and is the recommended way of working with a Kinvey instance. If you need to, you can still create one or more
Client instances of your own instead.
import Kinvey // Initialize each Kinvey instance let k1 = Client( appKey: "<#Instance 1 App Key#>", appSecret: "<#Instance 1 App Secret#>" ) let k2 = Client( appKey: "<#Instance 2 App Key#>", appSecret: "<#Instance 2 App Secret#>" ) // Then access data store collections separately on each instance let bookStoreK1 = try DataStore<Book>.collection(options: Options(client: k1)) let bookStoreK2 = try DataStore<Book>.collection(options: Options(client: k2))
instanceIdto the ID of their dedicated Kinvey instance.
You can find your Instance ID on the dashboard of the Kinvey Console, next to your App Key and App Secret.
Kinvey.sharedClient.initialize( appKey: "<#Your App Key#>", appSecret: "<#Your App Secret#>", instanceId: "<#Your Instance ID#>" ) { switch $0 { case .success(let user): if let user = user { print("User: \(user)") } case .failure(let error): print("Error: \(error)") } }
Verify Set Up
You can use
Client.ping() to verify that the app credentials were entered correctly. This function will contact the backend and verify that the library can communicate with your app. This code can be placed in your
UIApplicationDelegate.application(_, didFinishLaunchingWithOptions:) method after the
Client.initialize(). This function is good for while-in-development sanity check, but we do not recommend using it as part of a production application. The ping method is a not a robust way to test that the service is available.
Kinvey.sharedClient.ping() { (result: Result<EnvironmentInfo, Swift.Error>) in switch result { case .success(let envInfo): print(envInfo) case .failure(let error): print(error) } }
Set Instance Options
Beside the mandatory
appKey and
appSecret, you can initialize the Kinvey instance with a number of options.
For example, you can specify a longer or shorter timeout for requests to the instance.
Kinvey.sharedClient.initialize( appKey: "<#Your App Key#>", appSecret: "<#Your App Secret#>" ) { switch $0 { case .success(let user): if let user = user { print("User: \(user)") } case .failure(let error): print("Error: \(error)") } } Kinvey.sharedClient.options = try Options(timeout: 90)
If you've set the
timeout option through the
Options structure for a single request or an entire collection instance, or even for the same Kinvey instance, it will override the value of
timeoutInterval.
See the API reference for the full list of options.
The Options Structure
Whenever you need to set a different option value other then the default value for a single request, a collection instance, or a Kinvey instance, you can set it in the
Options structure. Although the SDK allows you to set some of the options separately, setting them through the structure is more convenient and tidy in most cases.
To see the complete list of values defined by the structure, check the API Reference.
All properties in the
Options structure are optional. The value of
nil is assumed for a property that you don't set explicitly, which is the default value.
To set options, first instantiate the
Options structure and then pass the instance to the request or collection or Kinvey instance. You can create as many
Options instances as you need and set their options differently.
Kinvey.sharedClient.initialize( appKey: "<#Your App Key#>", appSecret: "<#Your App Secret#>" ) { switch $0 { case .success(let user): if let user = user { print("User: \(user)") } case .failure(let error): print("Error: \(error)") } } Kinvey.sharedClient.options = try Options(timeout: 120)
The following example sets client app version and custom request properties which correspond to setting the custom headers
X-Kinvey-Client-App-Version and
X-Kinvey-Custom-Request-Properties respectively.
The
clientAppVersion option sets a version for your app that is sent with every network request. Is allows you to implement different behavior for the various app versions that you have deployed.
The
customRequestProperties options allows you to pass arbitrary values with your request, which you can then evaluate with business logic and act depending on the value.
let dataStore = try DataStore<Book>.collection() let options = try Options( clientAppVersion: "1.0.0", customRequestProperties: [ "myCustomString" : "test", "myCustomNumber" : 1 ] ) dataStore.save(book, options: options) { switch $0 { case .success(let book): print(book) case .failure(let error): print(error) } }. | https://devcenter.kinvey.com/ios/guides/getting-started | CC-MAIN-2019-47 | refinedweb | 1,074 | 57.47 |
Python alternatives for PHP functions
import re
def strspn(str1, str2, start=0, length=None):
if not length: length = len(str1)
return len(re.search('^[' + str2 + ']*', str1[start:start + length]).group(0))
(PHP 4, PHP 5)
strspn — Find length of initial segment matching mask".
The first string.
The second string.
The start position of the string to examine.
Negative value counts position from the end of a string.
The length of the string to examine.
Negative value sets length from the end of a string.
Returns the length of the initial segment of str1
which consists entirely of characters in str2
.
Example #1 strspn() example
<?phpecho strspn("foo", "o", 1, 2); // 2?>
Note: This function is
binary-safe. | http://www.php2python.com/wiki/function.strspn/ | CC-MAIN-2019-35 | refinedweb | 118 | 71.61 |
In this section, we will learn more about methods for selecting multiple rows and columns from a dataset read into pandas. We will also introduce some of the pandas data selection methods, and we will apply these methods to our real dataset to demonstrate the selection of a subset of data.
Let's get started by importing pandas and reading the data from zillow.com in the same manner as we did in the previous section. This is done as follows:
import pandas as pd zillow = pd.read_table('data-zillow.csv', sep=',') zillow.head()
The following is the output:
Next, let's look at some techniques ... | https://www.oreilly.com/library/view/mastering-exploratory-analysis/9781789619638/3ba3432a-7ad5-4f79-a1cb-c0209ad6d3d9.xhtml | CC-MAIN-2020-05 | refinedweb | 106 | 65.73 |
Operate on timestamps with Flux
Every point stored in InfluxDB has an associated timestamp. Use Flux to process and operate on timestamps to suit your needs.
- Convert timestamp format
- Calculate the duration between two timestamps
- Retrieve the current time
- Normalize irregular timestamps
- Use timestamps and durations together
If you’re just getting started with Flux queries, check out the following:
- Get started with Flux for a conceptual overview of Flux and parts of a Flux query.
- Execute queries to discover a variety of ways to run your queries.
Convert timestamp format
Unix nanosecond to RFC3339
Use the
time() function
to convert a Unix nanosecond timestamp
to an RFC3339 timestamp.
time(v: 1568808000000000000) // Returns 2019-09-18T12:00:00.000000000Z
RFC3339 to Unix nanosecond
Use the
uint() function
to convert an RFC3339 timestamp to a Unix nanosecond timestamp.
uint(v: 2019-09-18T12:00:00.000000000Z) // Returns 1568808000000000000
Calculate the duration between two timestamps
Flux doesn’t support duration column types.
To store a duration in a column, use the
string() function
to convert the duration to a string.
Retrieve the current time
Current UTC time
Use the
now() function to
return the current UTC time in RFC3339 format.
now()
now() is cached at runtime, so all instances of
now() in a Flux script
return the same value.
Current system time
Import the
system package and use the
system.time() function
to return the current system time of the host machine in RFC3339 format.
import "system" system.time()
system.time() returns the time it is executed, so each instance of
system.time()
in a Flux script returns a unique value.
Normalize irregular timestamps
To normalize irregular timestamps, truncate all
_time values to a specified unit
with the
truncateTimeColumn() function.
This is useful in
join()
and
pivot()
operations where points should align by time, but timestamps vary slightly.
data |> truncateTimeColumn(unit: 1m)
Input:
Output:
Use timestamps and durations together
Add a duration to a timestamp
date.add()
adds a duration to a specified time and returns the resulting time.
import "date" date.add(d: 6h, to: 2019-09-16T12:00:00Z) // Returns 2019-09-16T18:00:00.000000000Z
Subtract a duration from a timestamp
date.sub()
subtracts a duration from a specified time and returns the resulting time.
import "date" date.sub(d: 6h, from: 2019-09-16T12:00:00Z) // Returns 2019-09-16T06:00:00.000000000Z
Shift a timestamp forward or backward
The timeShift() function adds the specified duration of time to each value in time columns (
_start,
_stop,
_time).
Shift forward in time:
from(bucket: "example-bucket") |> range(start: -5m) |> timeShift(duration: 12h)
Shift backward in time:
from(bucket: "example-bucket") |> range(start: -5m) |> timeShift(duration: -12. | https://docs.influxdata.com/influxdb/v2.4/query-data/flux/operate-on-timestamps/ | CC-MAIN-2022-40 | refinedweb | 447 | 57.16 |
libs/libkipi/src
#include <configwidget.h>
Detailed Description
Definition at line 45 of file configwidget.h.
Constructor & Destructor Documentation
Default constructor.
Definition at line 104 of file configwidget.cpp.
Definition at line 141 of file configwidget.cpp.
Member Function Documentation
Return the number of plugins actived in the list.
Definition at line 196 of file configwidget.cpp.
Apply all changes about plugins selected to be hosted in KIPI host application.
Definition at line 146 of file configwidget.cpp.
Clear all selected plugins in the list.
Definition at line 183 of file configwidget.cpp.
Return the total number of plugins in the list.
Definition at line 191 of file configwidget.cpp.
Return the current string used to filter the plugins list.
Definition at line 243 of file configwidget.cpp.
Select all plugins in the list.
Reimplemented from QTreeView.
Definition at line 175 of file configwidget.cpp.
Set the string used to filter the plugins list.
signalSearchResult() is emitted when all is done.
Definition at line 222 of file configwidget.cpp.
Signal emitted when filtering is done through slotSetFilter().
Number of plugins found is sent when item relevant of filtering match the query.
Return the number of visible plugins in the list.
Definition at line 209 of file configwidget.cpp.
The documentation for this class was generated from the following files:
Documentation copyright © 1996-2020 The KDE developers.
Generated on Sun May 24 2020 23:25:52 by doxygen 1.8.7 written by Dimitri van Heesch, © 1997-2006
KDE's Doxygen guidelines are available online. | https://api.kde.org/4.x-api/kdegraphics-apidocs/libs/libkipi/src/html/classKIPI_1_1ConfigWidget.html | CC-MAIN-2020-24 | refinedweb | 254 | 54.59 |
C++/Loops
Loops allow a programmer to execute the same block of code repeatedly. We will make heavy use of conditional statements in this section.
The while Loop[edit | edit source]
The while loop is really the only necessary repetition construct. The for loop, coming up, can be duplicated using a while loop, and with more control. A simple negation can perform the same behavior as an until loop.
The syntax is as follows:
while ( ''condition'' ) { //body }
Again, the curly braces surrounding the body of the while loop indicate that multiple statements will be executed as part of this loop. If the actual code looked something like this:
while ( x == 4 ) y += x; x += 1;
means
while ( x == 4 ) { y += x; } x += 1;
There would be a problem. According to what was written, even though the second line after the while was indented, only the first line corresponds to the while loop. This is a huge problem because the variable involved in the condition (x) does not change, so it will always evaluate to true, making this an infinite loop. This could be alleviated by containing all statements intended to be a part of the loop body in { }.
while ( x == 4 ) { y += x; x += 1; }
The do...while Loop[edit | edit source]
The do...while loop is nearly identical to the while loop, but instead of checking the conditional statement before the loop starts, the do...while loop checks the conditional statement after the first run, then continuing onto another iteration.
The syntax is as follows:
do { //body } while (''condition'');
As you can see, it will run the loop at least once before checking the conditional.
The do...while loop is still haunted by infinite loops, so exercise the same caution with the do...while loop as you would with the while loop. Its usefulness is much more limited than the while loop, so use this only when necessary.
The for Loop[edit | edit source]
The
for loop is a loop that lets a programmer control exactly how many times a loop will iterate.
The syntax is as follows:
for (expression for initialization ; expression for testing ; expression for updating) { //body }
Here is an example of how the
for loop works:
for ( int i = 0; i < 10; ++i ) { std::cout << i+1 << std::endl; }
The code above and below are more or less equivalent.
int i = 0; while ( i < 10 ) { std::cout << i+1 << std::endl; ++i; }
What does this loop do? Prior to the first iteration, it sets the value of
i to
0. Next, it tests (like a normal while loop) if
i is less than 10. If the statement returns
true, the body of the loop is run and the program will print the value returned by the simple arithmetic statement
i+1 and move the terminal cursor down to the next line. After the loop is finished,
i is incremented (by 1), as specified in the update statement, and the conditional is tested again.
So, this loop will run a total of 10 times, printing the "i+1" each time. You've taught your program to count!
The variable used in for loops is generally an integer variable named i, j, or k, and is often initialized prior to the beginning of the for loop. Another option is to initialize the variable at the same time that you declare the variable's initial state:
for (int i = 0; i < 10; i++) { std::cout << i+1 << std::endl; }
If this is done properly, it is possible to nest for loops.
#include <iostream> using namespace std; int main(){ int input; cout << "How many missiles will you fire?" << endl; cin >> input; cout << "\n"; for (int i = 0; i < input; i++) { for (int j = 10; j > 0; j--) { cout << j << " "; } cout << "Missile " << i+1 << " has launched." << endl; } cout << "All missiles have been launched." << endl; return 0; }
What does this program do? The user is prompted to choose an integer, which is used in the conditional statement of the first for loop. Each iteration of the i loop will cause the j loop to run, which counts down from 10 to 1 before allowing the i loop to continue. The output will look something like this:
How many missiles will you fire? 3 10 9 8 7 6 5 4 3 2 1 Missile 1 has launched. 10 9 8 7 6 5 4 3 2 1 Missile 2 has launched. 10 9 8 7 6 5 4 3 2 1 Missile 3 has launched. All missiles have been launched.
Equivalence of C++ Looping Structures[edit | edit source]
The while loop can take the place of do...while loops if the condition for the while loop is "rigged" to be true at least the first time around.
The while loop can take the place of until loops by negating the condition that would be specified for the until loop, as explained above.
For and While loops[edit | edit source]
A for loop is structured as follows:
for (<initial expression> ; <condition> ; <update expression>) { <block of code> }
This can easily be reformatted as (do recognize the extra enclosing brackets, and the two extra semicolons after the expressions in order to turn them into statements):
{ <initial expression> ; while (<condition>) { <block of code> <update expression> ; } }
A for loop is more often used in by C++ programmers due to its conciseness as well as its separation of the looping logic (often using a loop control variable like "int i" or another simple iterator) from the loop's content. A while loop is often preferred if the initial statement or update statement requires more complex code than fits neatly into the for construct. However, the two are fully equivalent therefore it is ultimately a coding style decision, not a technical decision whether to use one or the other.
Infinite Loops[edit | edit source]
One common programming mistake is to create an infinite loop. An infinite loop refers to a loop, which under certain valid (or at least plausible) input, will never exit.
Beginning programmers should be careful to examine all the possible inputs into a loop to ensure that for each such set of inputs, there is an exit condition that will eventually be reached. In addition to preventing the program from hanging (i.e. never finishing), understanding all potential inputs and exit conditions demonstrates a solid understanding of the algorithm being written. Academic studies of programming languages often test students' ability to identify infinite loops for the very reason that doing so often requires a solid understanding of the underlying code.
Compilers, debuggers, and other programming tools can only help the programmer so far in detecting infinite loops. In the fully general case, it is not possible to automatically detect an infinite loop. This is known as the halting problem. While the halting problem is not solvable in the fully general case, it is possible to determine whether a loop will halt for some specific cases.
Example of Infinite loop[edit | edit source]
#include <iostream> int main() { while(1) { // printf("Infinite Loop\n"); std::cout << ("Infinite loop\n"); } return 0; }
This code will print infinite loop with out stopping. | https://en.wikiversity.org/wiki/C%2B%2B/Loops | CC-MAIN-2021-43 | refinedweb | 1,192 | 67.79 |
Talk:Osmarender
Contents
- 1 grouped SVG symbols and relative co-ordinates
- 2 segments and ways with no keys
- 3 logo on small maps
- 4 How do Iget Osmarender to work (Also Problems)
- 5 Projection
- 6 Rendering several .osm files into one SVG
- 7 rendering bug: oneway arrows
- 8 buildings
- 9 rendering of links
- 10 Problem with rendering of paths fixed
- 11 Ideas for rendering oneways
- 12 How do I get svg tiny?
- 13 Minor problem with rendering of residential roads
- 14 Problems with osmarender4
- 15 U-formed street name rendering
- 16 Issues with clipping and duplicates
- 17 Borders of water areas
- 18 Zoomlevel 18 for Osmarender (20 m/100 ft)
- 19 RDF description of maps created with Osmarender
- 20 Osmarender and Great Salt Lake
- 21 SVG problem: Invalid CSS selection
- 22 adding static svg-code via rules-file
grouped SVG symbols and relative co-ordinates
I'm trying to work out how to tweak it to be able to construct compound symbols for nodes, (for example. a church might look like a circle with a cross on top of it), a car park might be a white P on a blue rounded rectangle background, but can't work out how to do relative co-ordinates.
- I don't know either, yet. I've been working on a method for marking one-way streets, etc. Using a <marker> element to define an arrow but need to figure out how to size it relative to the width of the line it is on. If I get it worked out I'll let you know. 80n 18:38, 5 Apr 2006 (UTC)
segments and ways with no keys
I'm sure that it used to have a default method of dealing with segments and ways without any key/value defined that matches something in osm-map-features.xml. This doesn't seem to work at present. It doesn't draw any casings or fill (though ways defined with a valid "abutters" key get the surrounding shading applied)
- A rule that has k="~" v="~" will match any segment or way that has no tags at all. In osm-map-features.xml there is a commented out debug rule that will render all untagged elements as a thin red lines:
<rule k="~" v="~"> <line class='debug'/> </rule>
- You can also use k="something" v="~" to match any element that has a key of something whatever it's value. This is useful for making a default rendering of user-defined values. 80n 13:21, 3 May 2006 (UTC)
logo on small maps
On small maps, the logo and attribution info take up a very large amount of space (and sometimes don't even all fit on). Might it be possible to compute a scaling factor for their size, based on the size of the map data downloaded? Gagravarr 16:38, 4 Jun 2006 (UTC)
- The next version will include a dynamically positioned logo top left that stays the same size and location (relative to the screen) whatever zooming or panning that is done. It uses javascript to do this but don't know what will happen if javascript is not enabled or not supported by the SVG viewer. 80n 19:11, 4 Jun 2006 (UTC)
How do Iget Osmarender to work (Also Problems)
I could not reproduce this on Linux:
- First Try Firefox:
(Could not load the Url). Waiting for long time. Also a wget dont work:
- Actually it no longer works (for the area given above), and there is hardly any downloading going on. Bruce89 19:33, 16 Jun 2006 (UTC)
- I enter the account details, and then the Firefox waiting symbol rotating all the times. Sven Anders 19:03, 16 Jun 2006 (UTC)
- I realise it could just be the fact the server is very slow, it might take 30 minutes for an area that size to download. Bruce89 19:33, 16 Jun 2006 (UTC)
- Second Try Xalan:
(used a osm File from JOSM)
sven@machine:~$ java -cp xalan-j_2_7_0/xalan.jar org.apache.xalan.xslt.Process -in osm-map-features.xml -out data.svg (Fehler befindet sich an unbekannter Stelle)XSLT-Fehler (javax.xml.transform.TransformerConfigurationException): getAssociatedStylesheets failed
- The standard version expects a file named data.osm, if it is called josm2.osm then that might be the problem... 80n 18:22, 15 Jun 2006 (UTC)
- No i renamed it before.This is not the Problem Sven Anders 19:03, 16 Jun 2006 (UTC)
- when giving debug messages type "LANG=C; ava -cp xalan-j_2_7_0/xalan.jar org.apache.xalan.xslt.Process -in osm-map-features.xml -out data.svg" instead since the message will be in English then. Erik Johansson 13:02, 1 Jul 2006 (UTC)
- Third try xsltproc.
xsltproc Osm-map-features.xml josm2.osm >map.svg compilation error: file osmarender.xsl line 7 element rules xsltParseStylesheetProcess : document is not a stylesheet compilation error: file Osm-map-features.xml line 10 element rules xsltParseStylesheetProcess : document is not a stylesheet
Need Help! Thank you Sven Anders 08:26, 15 Jun 2006 (UTC)
Projection
It would be good if we could suggest a formula for working out the best projection to use (probably based on the latitude, and some sort of trig function). Gagravarr 08:48, 29 Aug 2006 (BST)
- The right way to do it would be to implement the Mercator projection algorithm () but that involves trig functions. XSL does not implement any trig functions natively. There are some XSL trig libraries but they all rely on proprietary extensions so don't work with all XSL processors.
- A simple hack (that has just occurred to me) would be to find a good projection value for each 5 degrees of latitute and then just include a small lookup table. It wouldn't be perfect but it would be an improvement. 80n 09:30, 29 Aug 2006 (BST)
Rendering several .osm files into one SVG
How hard would it be to make osmarender accept multiple .osm files, which'd go into a single svg?
(I'm thinking of taking the atom feed of pubs from an OpenGuide, passing it through a XSLT to turn it into a .osm file of nodes tagged as pubs, then having this + the normal osm data rendered together onto a single svg). --Gagravarr 20:15, 4 October 2006 (BST)
- It should be totally easy. Osmarender makes no assumptions about the ordering of nodes, segments or ways within a source .osm file. Chop out the <osm> and </osm> tags from each file, concatenate them and wrap in a new <osm> tag. That's the non-XML method ;-)
- I suspect that there could also be a relatively simple patch to osmarender.xsl that would process a .xml file containing a list of .osm files, but I haven't figured out what that would be yet. 80n 22:15, 4 October 2006 (BST)
rendering bug: oneway arrows
Does osmarender render arrows at the end of oneway ways? I thought this worked previously but don't see it tonight. Has k=oneway v=yes changed? Rw 03:42, 26 October 2006 (BST)
- It still works. Howver, the one-way arrows are tiny and you may not see them unless you zoom in a lot. Also its possible that they can be obscured at the end by the connecting road, especially if it is wider than the size of the arrow (a change to the render order in the rules file might fix that). A larger and clearer arrow is required and possibly some tweaking of the way it is positioned. 80n 07:15, 26 October 2006 (BST)
- I noticed the same type of problem. Indeed, sometimes they get obscured but it seems that the arrow is rendered at the end of a way, not in the middle of the vector. You can see it in this map of Leuven (it's not very clear, see the motorway_link at the top left of the image, the arrow is close to the trunk and not in the middle) Cimm 3 November 2006
- It doesn't work for me and the reason is a missing xml namespace for the marker elements. See . Jochen Topf 12 December 2006
buildings
Can we have support for building = ~ (wildcard, any type of building) in the default osmarender, so new types of building show up immediately rather than waiting for map_features and osmarender to be updated? Ojw 13:01, 6 November 2006 (UTC)
rendering of links
Currently motorway|trunk|primary_links are rendered together with the motorway|trunk|primary ways. I propose rendering the links before the main ways. See these pictures for the problem:
I think the order of drawing should be: primary_link, trunk_link, motorway_link, primary, trunk, motorway
Jochen Topf 9 December 2006
Problem with rendering of paths fixed
osmarender uses paths included with <use xlink...> and sometimes they don't render right. This is a problem for instance with rail lines which should get dotted lines. The reason for that is that the styles are attached to the surrounding <g> element and not the <use> element. explains how this work: The attributes from the <use> elements are copied to a newly created <g> element, NOT the parent <g> element.
The fix for this is pretty easy: Just copy the instructions attribute to the <use> element: <use xlink:<xsl:apply-templates</use>
Jochen Topf - 9 December 2006
Ideas for rendering oneways
The current rendering of one ways is suboptimal. Using an arrow at the beginning or end auf a way is not enough, there should be several arrows spaced along the path.
I experimented a bit:
- I have tried using several lines drawn over each other with different stroke-dashed styles to simulate an arrow appearing every so often. This works but looks rather pixelized.
- An older working draft of the SVG specification allows for any kind of vector graphic to be drawn along a path, but the current spec only allows text to be drawn there. So that doesn't work.
- I tried using unicode char U+25b6 (forward pointing triangle) several times in a text with spaces in between and drawn this text along a path. This works and doesn't look too bad. But it isn't clear to me how to best fit this into the osmarender xslt stuff.
This shows the unicode arrow stuff:
Creates with this svg: <text text-
<textPath xlink:href="#way_4048287" startOffset="0%"</textPath> </text>
Instead of having all those unicode chars they should be referenced with tref (see ), but I couldn't get that to work with inkscape.
Any other ideas?
-- Jochen Topf - 9 December 2006
How do I get svg tiny?
Hello everybody
I would really like to use the maps on my mobile phone, so I was thinking of tweeking osmarender.xsl in order to make it produce an SVG-TINY file instead of SVG-FULL.
I know I can use Adobe Illustrator to make the conversion from svg-full to svg-tiny. However I would really like to get rid of this extra conversion step.
Since tweeking osmarender.xsl seems a bit overwhelming, I would appreciate any advice (or any existent xsl code) which could help me automatically do the osm->svg-tiny conversion.
Many thanks in advance
Minor problem with rendering of residential roads
When a residential road ends in nothing the grey area around it is just cut straight off, but when it ends on another road it ends with a half-circle that stretches across the connected road. I think this should really be the other way around. The following image shows an example of how it looks.
In this particular case there are actually no houses in the road ends but that is another story. Karlskoging1 22:34, 19 February 2007 (UTC)
- An other problem with this gray areas is, that not always there are residents at both sides:
- In this example the residents of course do not live inside the river, but as the river has tag layer=-1 (which makes sense, because it flows below all those bridges) it vanishes next to residential roads. --Kumakyoo 15:50, 31 May 2007 (BST)
Problems with osmarender4
The current (SVN) version of osmarender4 seems to have some problems: The SVG created does not render correctly in Firefox (only a few features like railway stations and tunnels seem to be rendered, streets are not). In Inkscape it renders, but the colours of primary and trunk roads have been exchanged. (Trunk roads should be green, shouldn't they?) Håkon 11:42, 24 February 2007 (UTC)
I believe I have a similar issue. When viewing [[1]] in Firefox, the view is set to osmarender, nothing is displayed at this zoom level of 17. Maplint seems to be switched on by default. When I manually deselect Maplint the map is displayed. A quick solution for me would be if I could force Maplint off. Anyone know how to do this? --Eoin 06:17, 30 January 2008 (UTC)
U-formed street name rendering
U-formed streets sometimes get the name rendered around the bow. The general behaviour is to render the street name at the center of the street. Not that bad, but I wonder if osmarender can learn a more "intelligent" behaviour. A good example can be seen at the Windeckstraße in Karlsruhe Oberreut.
--SlowRider 15:22, 28 February 2007 (UTC)
Issues with clipping and duplicates
So I've been playing around with Osmarender today, and while I'm generally very impressed, I've run into two significant problems.
- Clipping doesn't seen to work very well. Having objects extend outside the selected area is still OK, but eg. in the case of Helsinki (a city on a peninsula, with ocean on west, south and east sides), the clipping of the ocean is terrible: the ocean objects are weirdly merged together, so the southern half of the city is submerged in blue, while the seas are white. See the attached picture (exported from Inkscape).
- Many objects, eg. the sea and descriptions of local buildings etc, have identical clones on top. Is there a reason for this?
For reference, I'm using an Osmarender version fresh from SVN, and here's how I generated my map:
wget mv map\?bbox\=24.907\,60.1519\,24.9795\,60.183 data.osm java -cp ../xalan-j_2_7_0/xalan.jar org.apache.xalan.xslt.Process -in osm-map-features-z17.xml -out map.svg
Any help would be appreciated. Jpatokal 14:09, 5 November 2007 (UTC)
- Thanks for the tip! So I downloaded Tiles@home from SVN, ran perl close-areas.pl 24.907 60.1519 24.9795 60.183 <../osmarender6/data.osm >../osmarender6/data2.osm (=same bounding box as in the original fetch) and then regenerated the SVG... but the end result is blank and absurdly huge (100x135 *meters* wide!) according to Inkscape, and painfully slow to render and absurdly huge according to Firefox. What am I doing wrong? How do the alternative "tilex/tiley" coordinates work? Jpatokal 16:09, 5 November 2007 (UTC)
Borders of water areas
It seems like all water areas, coast lines etc have a border in the same color at the water body. This makes thin land features that extends into the water appear even thinner example.
There are two solutions to this:
- Reduce width of border or remove it completely or
- Change colour of border. That makes users see that there actually is a border. I suggest a darker blue shade.
--Bengibollen 20:43, 16 December 2007 (UTC)
Zoomlevel 18 for Osmarender (20 m/100 ft)
Hi folks, if I'm looking at places like: I'm wondering if there is a plan to introduce an additional zoom level to osmarender? Cheers --Morray 15:25, 27 June 2009 (UTC)
- I would also like such a thing. Cyclemap,mapnik and noname has 20 m/100 ft zoom level even if cyclemap replaces (in most places) such zoom level tiles with blank white tiles. Then anyone who wants to render such tiles could enable any extra options in osmarender.
- I guess anyone who really wants this to be done(me, you, anyone) needs to devote lots of time for that, or ask someone else to devote lots of time for it. (hopefully someone who thinks it's a neat idea) Logictheo 08:51, 8 August 2009 (UTC)
- The source code is free to get I think. If I, you or anyone else is interested we just change it so it includes zoom level 18, right? logictheo 15:02, 4 September 2010 (BST)
- I added a ticket here: . If you want, you may comment. logictheo 16:30, 4 September 2010 (BST)
- I couldn't right now find where it was said, but the admins think the amount of additional data to be uploaded, stored and served would seriously slow down the currently available zoom levels. When disk/memory/network i/o is the worst bottleneck, multiplying the data would be unfeasible. Alv 22:12, 4 September 2010 (BST)
- Reading at Slippy_map_tilenames#Zoom_levels I see that Z18 "68.719.476.736" tiles is much more(~x4) compared to Z17 "17.179.869.184" tiles. Hmm, so I guess the only way would be to make a separate decentralised project for zoom level 18 that uses osmarender(data stored on clients computers). Phantasy name "TileZ18@home". I just noted down a project that even I might not be able to handle. Those interested in Z18 could get involved maybe(that inlcudes me, duh). That would be ideal. :) logictheo 13:51,13:57, 12 September 2010 (BST)
RDF description of maps created with Osmarender
While developing my label placement optimization tool OSMLabelOptimizer I noticed that it's not trivial to reconstruct the mapping information (such as which label belongs to which feature, which streets are connected or even which svg element draws the street) from the svg file. Finally I managed to do this, but it would be much easier if some more information about the map is included in the svg file. My idea is to do this with RDF (). Using this in combination with we could describe the whole svg map with all available OSM data and additional information. For example we could declare "the svg element with id 'test' draws the way with id '123'" where the last part could link to the linkedgeodata db. Additionally we could add something like "the svg text element with id 'test1' is label of the svg element with id 'test2'". I haven't yet thought about the details, but I think it would be possible to describe the svg map so that other programs could easily do cool things with it (displaying a list of all POI in the svg file, searching for a street name, displaying routes, etc.). What do you think about this? If you think this makes sense for Osmarender, I'd like to help you implement this. I just need a clue where I can put the code so that it is able to create the RDF in parallel to the svg map. -- Esscue 20:58, 23 August 2009 (UTC)
- I'm not sure if anybody is interested, but for completeness I would like to link to my user page where I'm going to develop a OWL ontology and OWL description of SVG maps generated with Osmarender: User:Esscue/OsmarenderOWL Have a look at it and send me an email with your opinion! -- Esscue 18:07, 20 September 2009 (UTC)
Osmarender and Great Salt Lake
Osmarender has had problems rendering the Great Salt Lake, Utah for several months. It doesn't draw too little water, but always too much. The excess water always fills the entire tile that is being rendered. If you would like to see it, it is at approximately at W 112.5 N 41. The lake is rather large and there are several tiles around the lake suffer from this problem. I have even made changes in the affected tiles to force rerendering an affected tile, but there is no effect on the apparent bug. The other renderers don't suffer from this problem.
Is this something that I've done to the data, or is this a bug in Osmarender? — Val42 04:55, 25 August 2009 (UTC)
SVG problem: Invalid CSS selection
Dear Osmarender experts,
I use Osmarender with XMLStarlet and get SVG files, which look nice at first glance. However, Adobe's SVG Viewer (3.0) messages "Invalid CSS selection" and Inkscape reports an internal error; both don't display the file. I already tried it using different SVG specifications from W3C but it did not work either. The operating system is Windows Vista. Have you already encountered this problem? Sorry that I can't give you further information up to now as I don't know where exactly the error occurred. --BlackEyedLion 11:43, 13 March 2010 (UTC)
- W3C's validator identifies the SVG file as SVG 1.1+IRI, XHTML5+ARIA plus MathML 2.0 (experimental) but does not validate the file as it is interrupted by an internal error. --BlackEyedLion 14:29, 14 March 2010 (UTC)
- I just started the conversion process anew - now with XMLSpy: I downloaded osm-map-features-z17.xml and osmarender.xsl, and I put a map file data.osm into the same folder. Then I opened osm-map-features-z17.xml in the XMLSpy and pressed F10. But in one of the first lines of osmarender.xsl two errors occurred:
- <msxsl:script this['node-set'] = function (x) { return x; } </msxsl:script> raised Syntax error and Identifier expected.
- Thank you for your help!
- (OS is Windows Vista, the map data are downloaded via the current API.)
--BlackEyedLion 17:27, 9 April 2010 (UTC)
adding static svg-code via rules-file
Is it possible to add static svg-code to the osmarender-output file? As far as I can see I can only add the given "templates" as drawing instructions like line, area, marker and so on. I would like to add some additional stuff (like a constant border or background). I am able to add svg-code inside of <defs>-tags - but how to do it outside? Jongleur 22:29, 26 May 2010 | http://wiki.openstreetmap.org/wiki/Talk:Osmarender | CC-MAIN-2016-44 | refinedweb | 3,697 | 70.84 |
private boolean isAbstract(Class klass) {
return (klass.getModifiers() & Modifier.ABSTRACT) > 0;
}
private boolean isFinal(Class klass) {
return (klass.getModifiers() & Modifier.FINAL) > 0;
}
What does this have to do with encapsulaton and minimal interfaces? In the code, they are odds with one another. The minimal interface is to simply have the getModifiers() method, but I believe this breaks encapsulation. I'm basically being exposed to how modifiers are implemented in the object, Class. Should I care that modifiers are really integers? If I got rid of the getModifiers() method and replaced it with corresponding methods isAbstract(), isFinal(), and etc. Then, I would not have a minimal interface, but I would have proper encapsulation.
I'm not trying to respark the humane vs. minimal interface debate. But, I'm trying to point out that sometimes they can be at odds with one another. And when they are at odds, which one would you pick? Me? I will always pick encapsulation because objects should always hide their internal implementations.
I also brought this topic to show that by placing more value on a minimal interface. You sacrifice one of the most important attributes of object-oriented programming: encapsulation. | http://blog.blainebuxton.com/2006/01/encapsulation-and-minimal-interfaces.html | CC-MAIN-2018-05 | refinedweb | 195 | 51.24 |
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Requirements: JDK 1.4. Read about more requirements on the
FreeTTS web site.
IBM's "Speech for Java"
Description: Implementation based on
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Requirements: JDK 1.1.7 or later or JDK 1.2 on Windows 95 with
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IBM's "Speech for Java" on Linux
Description: Beta version of "Speech for
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Requirements: RedHat Linux 6.0 with 32MB, and Blackdown JDK 1.1.7
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Description: Implementation for use with
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(some of which can be downloaded from Microsoft's web site).
Lernout & Hauspie's TTS for Java Speech
API
Description: Implementations based upon
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More information: Contact Edmund
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Conversa Web 3.0
Description: Conversa Web is a
voice-enabled Web browser that provides a range of facilities for
voice-navigation of the web by speech recognition and text-to-speech. The
developers of Conversa Web chose to write a JSAPI implementation for the speech
support.
Requirements: Windows 95/98 or NT 4.0 running on Intel Pentium 166
MHz processor or faster (or equivalent). Minimum of 32 MB RAM (64 MB
recommended). Multimedia system: sound card and speakers. Microsoft Internet
Explorer 4.0 or higher.
Festival
Description: Festival is a general
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Technology Research at the University of Edinburgh. It offers a full text to
speech system with various APIs, as well an environment for development and
research of speech synthesis techniques. It is written in C++ with a
Scheme-based command interpreter for general control and provides a binding to
the Java Speech API. Supports the English (British and American), Spanish and
Welsh languages.
Requirements: Festival runs on Suns (SunOS and Solaris), FreeBSD,
Linux, SGIs, HPs and DEC Alphas and is portable to other Unix machines.
Preliminary support is available for Windows 95 and NT. For details and
requirements see the Festival download page.
Elan Speech Cube
Description: Elan Speech Cube is a
Multilingual, multichannel, cross-operating system text-to-speech software
component for client-server architecture. Speech Cube is available with 2 TTS
technologies (Elan Tempo : diphone concatenation and Elan Sayso : unit
selection), covering 11 languages. Speech Cube native Java client supports
JSAPI/JSML.
Requirements: JDK 1.3 or later on Windows NT/2000/XP, Linux or
Solaris 2.7/2.8, Speech Cube V4.2 and higher.
About Elan Speech: Elan Speech is an established worldwide provider
of text-to-speech technology (TTS). Elan TTS transforms any IT generated text
into speech and reads it out loud.
Question:
Can I use HttpURLConnection class to make HTTPS requests in java 1.2?
Answer:
No, you need to use JSSE which includes HTTPS support - the ability to access
data such as HTML pages using HTTPS.Since Java 1.4 it is the part of J2SE.
Just take HttpsURLConnection class.!
Question:
Why does the InputStreamReader class has a read() method
that returns an int and not a char? Why we need additional steps for
converting?
Answer:
When end of stream is reached we always get the value -1. It is not possible
with char. How do we know when stream is empty? We can not use EOF, EOL and so
on...
in http proxy server i am
communicate with the webservers using URL and URLConnection. it is work fine.
but i didn't know how to access the mails through this.this
thread is
here...
hi all... i haf this problem...
when i add these codes to my program, it cannot read the next text files...
can help me solve it? this
thread is
here...
this
thread is
here...
Hi! I've created an
application with several frames. The main frame has a Menu Bar with all the
function that you can use, and each window has a Popup Menu with the most
useful functions. I thought (not enought, obviously) that you can create a
JMenuItem and use it in several Menus. For example, Create a 'Copy'
JMenuItem and use it in the Menu Bar and the JPopupMenu ... but it doesn't
work! When I add a JMenuItem, which was previously added to the Menu Bar, to
a JPopupMenu ... it disappear from de Menu Bar! Why? public class MainFrame extends JFrame{ this
thread is
here...
I have an Applet and a Servlet
changing java objects derilized multiple times. After a few seconds the applet
is started it stops. The applet is very easy, it has a thread that send and
object and recieve another one and the servlet just the opposite. Anyone can
help me?this
thread is
here...
I have a question when I use
InputEvent in AWT. Several constants are defined in InputEvent, such as
CTRL_MASK, SHIFT_MASK and META_MASK. Would you tell me what the META_MASK means?
Thanks!this
thread is
here...
Hi, I have a Frame_Grabber.dll
written in C. How can I use this DLL in my java application? I know
how to use java native interface, and to create my own dll, but what if the dll
already exists? Should I create my own DLL in any case? Or I can use this
FRAME_GRABBER.DLL directly and how pls?? Does someone know it?this
thread is
here...
Hi, I am arranging
components in gridbag layout. My parent container is JTabbedpane. But
JTextFields (and others too) are getting resized depending on there contents.
Can anyone help me in solving this problem. Thanks in advance,
Shaileshthis
thread is | http://javafaq.nu/java-article355.html | CC-MAIN-2018-13 | refinedweb | 1,181 | 69.68 |
, co-workers or students, consider building a package. The chances of broken dependencies and ease of installing everything outweighs the effort of learning how to build one. If you feel your functions (that may be new in some respect) could benefit an even wider audience, consider submitting it to CRAN (I will not discuss how to do that here, but do read the Ripley reference I mention later).
I have set out to build a test package to prepare myself when the time comes and will really need to build one of my own. This here is an attempt I made to document steps I took when building a dummy package (called texasranger (yes, THE Texas Ranger!)) with one single function. I have attempted to build documentation and all other ladeeda things that are mandatory for the package to check out O.K. when building it.
Before you dig into the actual preparation and building itself, you will need a bunch of tools. These come in a bag with a linux distribution, but you will have to add them yourself if you’re on Windows. This is basically the only thing that is different when trying to build a package on Windows/Linux. I will not go into details regarding these tools (perl, MS html help compiler, if you have C/C++/FORTRAN code you will need GNU compiler set) , a TeX distro), – I will, however, advise you to check out Making R package under Windows (P. Rossi). There, you will find a detailed description (see page 2-6) of how to proceed to get all the correct tools and how to set them up. When you have done so, you are invited to come back here. Feel free to follow just mentioned tutorial, as it goes a bit more in-depth with explaining various aspects. The author warns that MiKTeX will not work (see the datum of the document), but things might have changed since then and it now works, at least for me.
I have followed the aforementioned Making R package under Windows (by P. Rossi), slides Making an R package made by R.M. Ripley and of course the now famous Writing R Extensions (WRE) by R dev core team (you are referred to this document everywhere). I would advise everyone to read them in this listed order – or at least read WRE last. First two can be read from cover to cover in a few minutes – the last one is a good reference document for those pesky “details”. In my experience, I started to appreciate WRE only after I have read the first two documents.
Enough chit-chat, let’s get cracking!
1. These are the paths I entered (see document by Rossi what this is all about) to enable all the tools so that I can access them from the Command prompt (Command prompt can be found under Accessories, another term for it may be Terminal or Console on different OSs):
c:/rtools/bin;c:/program files/miktex 2.8/miktex/bin;c:/program files/ghostgum/gsview;C:/strawberry/perl/bin;c:/program files/r/r-2.11.0/bin;c:/program files/help workshop%SystemRoot%\system32;%SystemRoot%;%SystemRoot%\System32\Wbem;%SYSTEMROOT%\System32\WindowsPowerShell\v1.0\;C:\strawberry\c\bin;C:\strawberry\perl\site\bin
2. Use R function
package.skeleton()
to create directory structure and some files (DESCRIPTION and NAMESPACE). I used the following arguments:
package.skeleton(name = "texasranger", list = c("bancaMancaHuman"), namespace = TRUE) #I only have one function, but you can list them more
See argument code_files for an alternative way of telling the function where to read your functions. I suspect this may be very handy if you have each function in a separate file.
3. Fill out DESCRIPTION and NAMESPACE (if you decide to have a name space, read more @ WRE document). Pay special attention to export, import, useDynLib… All of the above mentioned documentation will help guide you through the process with minimal effort.
A side (but important) note. You should write your functions without them calling
require()
or
source()
to dig up other function and packages. Read more about NAMESPACE and how to specify which functions and packages to “export” (or “import”) and how.
4. Create documentation files. This is said to be the trickiest part. I still don’t have much experience with this so I can’t judge how tricky it can be – but I can tell you that it may be time consuming. Make sure you take time to document your functions well. If you were smart, you wrote all this down while you were writing (or preparing to write) a function and this should be more or less a session of copy-paste. Use
prompt(function, file = "filename.Rd")
to create template Rd files ready for editing. They are more or less self explanatory (with plenty of instructions). It help if you know LaTeX, but not necessary. Also, I suspect the function may dump the files into the correct /man directory automatically – if it doesn’t, do give it a hand and move the files there yourself. Perhaps worth mentioning is that if you want to reference to functions outside your package, use(notice the options square brackets [])
\code{\link[package:function]{function}}
, e.g.
\code{\link[raster:polygonsToRaster]{polygonsToRaster}}
or
\code{\link[utils:remove.packages]{remove_packages}}
- To refer to “internal” package function (those visible by the user), use
\code{\link{function_name}}
4a. If you have datasets you wish to include in your package (assuming those in library(help=”datasets”) are not sufficient), you will need to do two things. First, prepare your object (list, data.frame, matrix…). Save it and prepare documentation. Saved .rda file goes to data/ directory. The documentation file goes into the same directory (man/) as other .Rd files. If your dataset is not bigger than 1 MB you shouldn’t worry, otherwise consult the Manual on how to prepare a
save(my.dataset, file = "my.dataset.rda") # move to data/ folder promptData(my.dataset, filename ="my.dataset.rda.Rd") # move to man/ folder¸and edit
4b. You should also build a vignette, where you can explain at greater length what your package is about and maybe give a detailed (or more detailed) workflow with the accompanying functions. You can use Sweave or knitr, and the folder to place your .Rnw file is vignettes/.
5. To check the documentation for errors, use
R CMD Rd2txt filename.Rd
and/or
R CMD Rdconv -t=html -o=filename.html filename.Rd
6. Next, you should run a check on your package before you build it. You should run it from the directory where the package directory is located. I’ve dumped my package contents to d:/workspace/texasranger/ and executed the commands from d:/workspace/
R CMD check
If you get any errors, you will be referred to the output file. READ and UNDERSTAND it.
7. Build the package with the command
R CMD build package_name
This will create a file and will add a version (as specified in the DESCRIPTION file, i.e. package_name_1.0-1.tar.gz, see WRE for specifics on package version directives).
package_name is actually the name of the directory (which should be the name of your package as well).
If you use Windows, you can build a .zip file AND install the package (uses install.packages) at the same time. Use command
R CMD INSTALL --build package_name_1.0-1.tar.gz
8. Rejoy... | http://www.r-bloggers.com/building-an-r-package-under-windows-without-c-c-or-fortran-code/ | CC-MAIN-2015-22 | refinedweb | 1,246 | 64.71 |
This topic describes how to use Spatial Sound in your Unity projects. It covers the required plugin files as well as the Unity components and properties that enable Spatial Sound.
Spatial Sound, in Unity, is enabled using an audio spatializer plugin. The plugin files are bundled directly into Unity so enabling spatial sound is as easy as going to Edit > Audio > Spatializer and enabling the Microsoft HRTF extension. Since Microsoft Spatial Sound only supports 48000 currently, you should also set your System Sample Rate to 48000 to prevent an HRTF failure in the rare case that your system output device is not set to 48000 already:
Your Unity project is now configured to use Spatial Sound.
Note that, while the Windows 10 SDK may be used to build HoloLens apps on Windows versions prior to Windows 10, if you aren't using windows 10, you will not get Spatial Sound in the editor nor on the device.
Spatial Sound is used in your Unity project by adjusting three settings on your Audio Source components. The following steps will configure your Audio Source components for Spatial Sound.
Your sounds now realistically exist inside your project's environment!
It is strongly recommended that you become familiar with the Spatial Sound design guidelines. These guidelines help to integrate your audio seamlessly into your project and to further immerse your users into the experience you have created.
The Microsoft Spatial Sound plugin provides additional parameters that can be set, on a per Audio Source basis, to allow additional control of the audio simulation. These parameters are the minimum and maximum gain, the unity gain distance and the size of the simulated room.
The minimum gain applied (from -96 to +12 decibels) at any distance. The default value is -96 decibels.
The maximum gain applied (from -96 to +12 decibels) at any distance. The default value is +12 decibels.
The distance, in meters (from 0.05 to infinity), at which the gain is 0 decibels. The default value is 1 meter.
The size of room that is being simulated by Spatial Sound. The approximate sizes of the rooms are; small (an office to a small conference room), medium (a large conference room) and large (an auditorium). You can also specify a room size of none to simulate an outdoor environment. The default room size is small.
The HoloToolkit for Unity provides a static class that makes setting the Spatial Sound settings easy. This class can be found in the HoloToolkit\SpatialSound folder and can be called from any script in your project. It is recommended that you set these parameters on each Audio Source component in your project. The following example shows selecting the medium room size for an attached Audio Source.
AudioSource audioSource = gameObject.GetComponent<AudioSource>() if (audioSource != null) { SpatialSoundSettings.SetRoomSize(audioSource, SpatialMappingRoomSizes.Medium); }
If you don't want to use the excellent Audio tools in HoloToolkit, here is how you would change HRTF Parameters. You can copy/paste this into a script called SetHRTF.cs that you will want to attach to every HRTF AudioSource. It lets you change parameters important to HRTF.
using UnityEngine; using System.Collections; public class SetHRTF : MonoBehaviour { public enum ROOMSIZE { Small, Medium, Large, None }; public ROOMSIZE room = ROOMSIZE.Small; // Small is regarded as the "most average" // defaults and docs from MSDN // public float mingain = -96f; // The minimum gain limit applied at any distance, from -96 to + 12 public float maxgain = 12f; // The maximum gain applied at any distance, from -96 to + 12 public float unityGainDistance = 1; // The distance at which the gain applied is 0dB, from 0.05 to infinity public float bypassCurves = 1; // if > 0, will bypass Unity's volume attenuation and make a more accurate volume simulation automatically in the plugin AudioSource audiosource; void Awake() { audiosource = this.gameObject.GetComponent<AudioSource>(); if (audiosource == null) { print("SetHRTFParams needs an audio source to do anything."); return; } audiosource.spatialize = 1; // we DO want spatialized audio audiosource.spread = 0; // we dont want to reduce our angle of hearing audiosource.spatialBlend = 1; // we do want to hear spatialized audio audiosource.SetSpatializerFloat(1, (float)room); // 1 is the roomsize param audiosource.SetSpatializerFloat(2, mingain); // 2 is the mingain param audiosource.SetSpatializerFloat(3, maxgain); // 3 is the maxgain param audiosource.SetSpatializerFloat(4, unityGainDistance); // 4 is the unitygain param audiosource.SetSpatializerFloat(5, bypassCurves ); // 5 is bypassCurves, which is usually a good idea } } | https://developer.microsoft.com/en-us/windows/holographic/spatial_sound_in_unity | CC-MAIN-2017-04 | refinedweb | 727 | 56.55 |
In this tutorial we will check how to generate a random integer using the micro:bit board and MicroPython.
Introduction
In this tutorial we will check how to generate a random integer using the micro:bit board and MicroPython.
To achieve this, we will use the random module from MicroPython, which offers some simple functions to get random results in our programs.
As already mentioned, we are going to check how to generate a random integer number, but this module offers other interesting functions.
The code
The code for this tutorial will be very simple. So, the first thing we will do is importing the random module, so we have access to the function we need to generate the random integer.
import random
Then, we simply need to call the randint function, in order to generate a random integer. As input, this function receives a start and an end integer parameters, which determine the range of values from which the random integer will be generated.
So, if we pass a start value of 0 and an end value of 10 (like in the code below), we will get a random number between 0 and 10. Note that the start and end values are included in the interval of possible numbers that will be generated [1].
print(random.randint(0,10))
Just to obtain some additional results, we will do a couple more calls to the randint function.
print(random.randint(0,10)) print(random.randint(0,10)) print(random.randint(0,10))
The final complete code can be seen below.
import random print(random.randint(0,10)) print(random.randint(0,10)) print(random.randint(0,10)) print(random.randint(0,10))
Testing the code
To test the code, run the previous script on the micro:bit using a tool of your choice. I’ll be using uPyCraft, a MicroPython IDE.
Upon running the code, you should get an output similar to figure 1. Naturally, the numbers you will obtain will probably be different from mine, due to the fact that they were generated randomly.
References
[1] | https://techtutorialsx.com/2019/03/24/microbit-micropython-generating-random-numbers/ | CC-MAIN-2019-43 | refinedweb | 347 | 52.9 |
Inside TFS
In Part 2 of his column, Mickey Gousset dives deeper into the Test Explorer window.
(This article applies to all editions of Visual Studio 2012, except Test Professional 2012)
In Part 1 of this column, I defined test-driven development (TDD), why it's important, why it's a successful development methodology, and the basics of how to implement TDD using the unit testing framework in Visual Studio 2012. In this column, I'm going to continue to build off the example code in Part 1, and continue to dive into the unit testing framework in Visual Studio 2012, specifically the different features provided by Test Explorer to help with the testing process.
When I left off last time, I'd created the stub code for the Add method and run the unit test for the method. The test failed because the code hadn't been implemented yet. This implies that I need to implement the actual method. I can add the following line of code to the method to implement the logic:
return (p1 + p2);
Open the Test Explorer window, and click the Run All link to run all the unit tests. The solution will compile and the tests will run. This time the test will pass, as expected.
Let's dive into the Test Explorer window in more detail. The Test Explorer window is my interface into the unit tests in my solution. The Test Explorer window, by default, groups my tests into multiple categories, such as Failed Tests, Not Run Tests and Passed Tests. This lets me easily see new tests that have been created but not yet executed, so I can run just those tests in a test run. To see this, add a new test method, SubtractTest, to the BasicMathFunctionsTest class, as shown in Figure 1.
As can be seen in Figure 1, a new method was added to the BasicMathFunctionsTest.cs class, called SubtractTest. If I compile the solution, the unit test framework will pick up on the new test method and add it to the Test Explorer window under the Not Run Tests category, as shown. This provides a quick and easy way to see what new tests are available.
Notice how the method currently contains one line of code, Assert.Inconclusive(). If I just create the test method without any code and run it, it would pass. While this is perfectly valid, it doesn't help me identify any test stub methods I might have created, i.e. test methods where I just created the shell of the method, with the intention of adding the actual test code later. Adding an Assert.Inconclusive() call to any unfinished test method will cause that test method to show up as skipped in the Test Explorer window (Figure 2). Any methods showing up as skipped would indicate test methods that need to be modified/created.
The Test Explorer window can be used to organize tests into Test Playlists. Right-click on a test, and select Add to Playlist | New Playlist to create a new playlist file. This creates an XML file with a .PLAYLIST extension, containing the names of all the tests in that playlist. I can now select the Playlist drop down box in the Test Explorer window and select the created playlist. This will display in Test Explorer only the tests contained in the playlist. To see all the tests available in a solution, click the Playlist drop down box again and select All Tests.
By default, the tests in Test Explorer are arranged by test outcome, such as passed, failed or skipped. You can also arrange them by test class, test duration or test project.
Another methodology starting to appear is the idea of continuous testing. The idea is that I should be continuously testing my code as I build it. Visual Studio 2012 supports this by allowing me to automatically run my unit tests as part of my personal build process. Instead of building my solution and then opening Test Explorer and running my tests, I can have the tests run automatically as soon as I build the solution. This saves time, as well as providing immediate feedback on the build process. To do this, click the Run Tests After Build button in the top left of the Test Explorer.
The past two columns have delved into how to perform TDD using Visual Studio 2012, and a more in-depth look into the Test Explorer window. Visual Studio 2012 provides many more testing opportunities for TDD, from the use of other testing frameworks to the use of shims and fakes to help mock out code to test. All of these tools can be used to help create better-tested code. | http://visualstudiomagazine.com/articles/2013/08/26/test-driven-development-vs2012.aspx | CC-MAIN-2014-52 | refinedweb | 791 | 70.13 |
public class LoginScreen? extends HTMLPage { InputField? nameField; public LoginScreen?() { Layout l = new FlowLayout(); l.add(new TextField("Name")); nameField = new InputField?() l.add(nameField); setLayout(l); } ... }The system should generate HTML from this for me. I can even create add a TreeView widget to the page, and the system will generate the appropriate HTML + JavaScript to make it all work. Maybe Swing is the wrong model. Maybe it should be more like SmalltalkLanguage. Or whatever GUI-building toolkit you prefer. But wouldn't this be better than dealing with HTML? There must be tools out there that let you work like this. What are they? ASP.NET provides such a programmatic interface and it works quite well. However, in practice, you'll still need to munge HTML if you want something that's complicated and pretty. ThereIsNothingPerlCannotDo. You want CGI.pm:
#!/usr/local/bin/perl use CGI qw(:standard); print header; print; } print end_html;This and much, much more at . | http://c2.com/cgi/wiki?WhyDoWeWriteHtml | CC-MAIN-2014-10 | refinedweb | 160 | 71.41 |
Details
Description
Currently the Java reflect API treats shorts as ints. This fails however to naturally handle shorts in arrays and other containers. It would be better if we used a distinct type for reflected Java short values.
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Floats aren't encoded varint, are they? I can't see the advantage there, the high bits will be set too frequently.
Yes, they're not encoded varint. Totally bogus statement on my part.
[Thoughts on a unified API]
Scott, I totally see where you're going. I think an "annotation-based" API would be a great addition. Even though it's possible to unify "specific" and "reflection" APIs under that one roof, I really like the simplicity of the "specific" API. This comes in part from some experience with Protocol Buffers. Yes, it sucks to wrap each of your data classes with another class, but, in practice, this is often easier to follow than stuff that does some magic behind your back.
In terms of examples, ORMs solve the same sorts of problems. Hiberate and friends deserve a look too, though I agree Jackson is more similar.
It also seems pretty arbitrary that integers, longs, and floats are represented with zigzag varint encoding, but shorts are always two bytes.
Floats aren't encoded varint, are they? I can't see the advantage there, the high bits will be set too frequently.
At this point I am not concerned about performance here.
Thats fair.
It wouldn't surprise me if Avro evolved to have int8, int16, int32, int64 and fixed8, fixed,16, fixed32, fixed64 "types"
Maybe, but if this happens, it sounds like Avro 2.0 not Avro 1.x.
Also,, there would be no benefit to varint of one byte, and for the int16 case there may be very little or no benefit. Its easy to speculate that a int32 is very often less than 2^20 in size. Its hard to speculate that shorts are mostly less than 2^6 and not frequently more than 2^13.
Not sure what you mean by mix-ins, but, yes, you could annotate the field in the class whose schema is being induced.
Basically if you don't want, or can't change class A, you can write MixIn class B that has annotations that "target" the methods and members of class A. See:
The goal, is to allow annotating a class you can't change the source code for.
Ok, if we're talking about the long term Reflect API, I will add this:
I have been starting to dig in to using Avro myself, and thinking about schema evolution. I don't particularly like the Specific API and its code generation, I'd generally rather direct a schema at my own classes for most use cases. I don't want to use Reflection either, with its restrictions and performance pitfalls (my requirements differ from those Doug is working on for Hadoop RPC significantly).
I think that these two APIs can be combined in one annotations based API. Sure, we can still have code generation from avro schemas with basic defaults to create classes, but that step can be optional, even for inter-language use cases.
Imagine something like this.
You have a pre-existing class, Stuff, and you want to define how it is serialized. You make an Avro schema for it, to share with other languages/tools. Now, you want to map the two together. Using Specific, you have to write wrapper code to read the Avro generated class into your current class (that has a little bit of logic in it, maybe a custom hashCode() and equals(), a few other constructors for test cases , and some setters and getters that aren't just "return foo" and "this.foo = foo". If this class is an already long lived class with lots of unit tests, there aren't a lot of nice ways to do this without refactoring more than just the class. More importantly, if you have 40 or so such classes —
Reflect can somewhat get around this, but then if you want to share the data with other languages and tools you've just exposed your Java implementation of your object to the world... I'd rather not have a schema change just because I changed some internal data structure already encapsulated with getters/setters.
Ideally, I would like to just annotate the class with something that says "this is serializeable with Avro with an avro type named org.something.X".
Then map the getters/setters or the fields to avro fields, and build any custom logic there if needed to deal with versions. Being able to map to a constructor would be cool too (like Jackson), but less important at the start.
We could even set it up to map projected schemas – "this class can be serialized as org.something.X, or the projection org.something.minimalX if method 'isMinimal()' returns true""
This same mapping can be done with an annotation MixIn if the class can't be modified at this time.
Now, when decoding anything where an avro tye of X is encountered, it just builds the object as instructed by the annotations. Of course, this can all be optimized early on at class loading time rather than with runtime reflection with something like ASM.
It may even be possible to just 'borrow' Jackson's annotations entirely, and be nearly or completely compatible with those.
The reason why I say that a 'complete' annotation style API can replace both reflection and specific, is that the rules for specific can be one set of defaults – what to construct when a type does not map to a known class, and the reflection default rules the other (how to serialize when a class is not annotated). The need to generate classes at compile time might go away (It can be defined when first encountered with ASM). The default behavior for both cases can be defined as some sort of Mix-In default: When reflecting, if you find a short, serialize it as an avro fixed 2 byte quantity. When generating an object from a type that is not declared, create org.apache.commons.MutableInteger for avro ints.
I had intended to create a ticket for something like the above after learning more and exploring – I should have more time over the next couple months to do more than observe and comment.
I like the new patch. +1.
Offline Philip suggested it might be better to map short to:{"type": "int", "java-class": "java.lang.Short"}
I like this approach better. Here's a version of the patch that implements it.
> BigEndian, it appears. Does that fact belong in documentation for the Reflect API?
Yes. I added this to ReflectData, along with the documentation of other mappings.
> On reading, it will be a bit faster if you avoid the array allocation
At this point I am not concerned about performance here. The reflect API already suffers from known performance problems. The correct handling of shorts is required for reflection to support existing Hadoop RPC protocols, a primary target for the reflect API. If this does show up as a performance bottleneck in benchmarks then we should address it.
> Just to confirm, the "fixed" type doesn't store a length byte, so there is no extra overhead.
That's correct.
> What about char? Its just an unsigned short.
If we need, we could add support for that in a separate issue.
It wouldn't surprise me if Avro evolved to have int8, int16, int32, int64 and fixed8, fixed,16, fixed32, fixed64 "types" (all implemented using varint and fixed, but with the APIs responsible for bounds checking) eventually. But that's neither here nor there.
Not sure what you mean by mix-ins, but, yes, you could annotate the field in the class whose schema is being induced.
I understand the expedience of modelling shorts this way. I worry that it will make the reflection API harder to evolve. It's saying that from now on, all Java shorts shall be 2-byte fixeds, and that's going to be the default forever more.
I haven't thought about the evolution required of the reflect API. Is it possible to change, later, how the same Java class is serialized?
> Could you just call it "Short"?
I'm hesitant to do this now. It seems tantamount to extending the specification with a new type. We could perhaps add it to a standard "lib" namespace or somesuch. But, until we've established that, an application that needs a two-byte quantity should define its schema to use an application-specific fixed type in its schemas.
The use of java.lang.Short highlights a place where folks are using a java-specific type. A mechanism for mapping arbitrary application-specific Avro types to unboxed Java types is not straightforward today. I don't want to solve that problem today. Today I want to make applications that use shorts work using reflection. And I don't mind that appearing as a wart on that application until we solve these harder problems.
> It also seems pretty arbitrary that integers, longs, and floats are represented with zigzag varint encoding, but shorts are always two bytes.
Yes, it is arbitrary. We could instead model shorts as a one-field int record. That would not restrict its range, and might use more space in some cases. Six of one.
> Have you considered using an approach similar to the annotation approach in
AVRO-242 for this?
Yes. What type would you annotate? Would you annotate each field that's a short? Would we implement mixin annotations?
Using "java.lang.Short" in a name seems suboptimal. Could you just call it "Short"? It seems like the schemas induced by reflection should ideally be language-neutral.
It also seems pretty arbitrary that integers, longs, and floats are represented with zigzag varint encoding, but shorts are always two bytes.
Have you considered using an approach similar to the annotation approach in
AVRO-242 for this? Even for something like longs, you're sometimes going to want to hint about what encoding to use (fixed vs. varint).
– Philip
BigEndian, it appears. Does that fact belong in documentation for the Reflect API?
On reading, it will be a bit faster if you avoid the array allocation – at least until scalar replacement or stack allocation works well in the JVM.
This will not show up in microbenchmarks, but can be rather significant in more complicated high throughput systems with enough other objects of 'medium' lifetimes. (The same is true of autoboxing and creating objects members instead of intrinsic members.)
If you can avoid an allocation and not affect performance or functionality, its a good idea to do so. This could be done with two one byte reads or by adding a ByteBuffer - like "readShort()" method to Decoder.
Just to confirm, the "fixed" type doesn't store a length byte, so there is no extra overhead.
What about char? Its just an unsigned short.
With this patch, reflection maps short values to the schema:{"type": "fixed", "name": "java.lang.Short", "size": 2}
.
Tests are added for arrays of shorts.
I just committed this. | https://issues.apache.org/jira/browse/AVRO-249?focusedCommentId=12786097&page=com.atlassian.jira.plugin.system.issuetabpanels:comment-tabpanel | CC-MAIN-2015-35 | refinedweb | 1,881 | 72.76 |
Similar to Binary Search, Jump Search or Block Search is an algorithm only for ordered (sorted) lists. The major idea behind this algorithm is to make less comparisons by skipping a definite amount of elements in between the elements getting compared leading to less time required for the searching process.
It can be classified as an improvement of the linear search algorithm since it depends on linear search to perform the actual comparison when searching for a value.
In Jump search, it is not necessary to scan all the elements in the list to find the desired value. We just check an element and if it is less than the desired value, then some of the elements following it are skipped by jumping ahead.
After moving a little forward again, the element is checked. If the checked element is greater than the desired value, then we have a boundary and we are sure that the desired value lies between the previously checked element and the currently checked element. However, if the checked element is less than the value being searched for, then we again make a small jump and repeat the process.
Given a sorted list, instead of searching through the list elements incrementally, we search by
jump. The optimal size for
jump is $\sqrt{N}$ where the
N is the length of the list.
So in our input list
list_of_numbers, if we have a jump size of
jump then our algorithm will consider the elements between
list_of_numbers[0] and
list_of_numbers[0 + number_of_jumps], if the key element is not found then we will consider the other elements between
list_of_numbers[0 + number_of_jumps] and
list_of_numbers[0 + 2*number_of_jumps], again if the key element is not found then we consider the elements between
list_of_numbers[0 + 2*number_of_jumps], and
list_of_numbers[0 + 3*number_of_jumps] and so on.
With each jump, we store the previous value we looked at and its index. When we find a set of values where
list_of_numbers[i] < key_element < list_of_numbers[i + number_of_jumps], we perform a linear search with
list_of_numbers[i] as the left-most element and
list_of_numbers[i + number_of_jumps] as the right-most element in our search block.
For example, consider a list of
[1, 2, 3, 4, 5, 6, 7, 8, 9].The length of the list is
9 and the size of
jump is
3. If we have to find the key element
8 then the following steps are performed using the Jump search technique.
Step 1: First three elements are checked. Since 3 is smaller than 8, we will have to make a jump ahead.
Step 2: Next three elements are checked. Since 6 is smaller than 8, we will have to make a jump ahead.
Step 3: Next three elements are checked. Since 9 is greater than 8, the desired value lies within the current boundary.
Step 4: A linear search is now done to find the value in the array.
import math def jump_search(list_of_numbers, key): list_length = len(list_of_numbers) # Calculate jump size jump = int(math.sqrt(list_length)) left, right = 0, 0 # These are the index values # Find the block where key element is present (if it is present) while left < list_length and list_of_numbers[left] <= key: right = min(list_length - 1, left + jump) if list_of_numbers[left] <= key <= list_of_numbers[right]: break left += jump if left >= list_length or list_of_numbers[left] > key: return -1 right = min(list_length - 1, right) i = left # Do linear search for key element in the block while i <= right and list_of_numbers[i] <= key: if list_of_numbers[i] == key: return i # Return the position of the key element i += 1 return -1 def user_input(): list_of_numbers = list() total_elements = input("Enter a list of numbers in ascending order with space between each other: ").split() for each_element in range(len(total_elements)): list_of_numbers.append(int(total_elements[each_element])) key = int(input("Enter the Key number to search: ")) index = jump_search(list_of_numbers, key) if index == -1: print("Entered key is not present") else: print(f"Key number {key} is found in Position {index + 1}") if __name__ == "__main__": user_input()
Enter a list of numbers in ascending order with space between each other: 1 2 3 4 5 6 7 8 9 Enter the Key number to search: 8 Key number 8 is found in Position 8
Elements -> 1 2 3 4 5 6 7 8 9 Index -> 0 1 2 3 4 5 6 7 8 left = 0 right = 0 list_length = 9 jump = 3 while 0 < 9 and 1 <= 8: right = min(8, 3) right is 3 if 1 <= 8 <= 4 condition is False left = 3 while 3 < 9 and 4 <= 8: right = min(8, 6) right is 6 if 4 <= 8 <= 7: condition is False left = 6 while 6 < 9 and 7 <= 8: right = min(8, 9) right is 8 if 7 <= 8 <= 9: condition is True break right = min(8, 8) right is 8 i = left i is 6 while 6 <= 8 and 7 <= 8: if 7 == 8: condition is False i += 1 i is 7 while 7 <= 8 and 8 <= 8: if 8 == 8: condition is True return 7 here 7 is the index position and not the actual element print -> i + 1 position | https://nbviewer.jupyter.org/github/gowrishankarnath/Dr.AIT_Python_Lab_2019/blob/master/Program_3.ipynb | CC-MAIN-2020-40 | refinedweb | 849 | 58.25 |
.sharedClient = [[SentryClient alloc] initWithDsn:@"your dsn" didFailWithError:nil]; [SentryClient.sharedClient startCrashHandlerWithError:nil]; return YES; }Instructions for Swift
import Sentry func application( application: UIApplication, didFinishLaunchingWithOptions launchOptions: [NSObject: AnyObject]?) -> Bool { Client.shared = try? Client(dsn: "your dsn") try? Client.shared?.startCrashHandler() return true }
Sentry automatically captures crashes recorded on macOS, iOS, and tvOS.
View the Sentry for Cocoa documentation for more information.
Get iOS error monitoring with complete stack traces
See details like function, filename, and line number so you never have to guess.
Reveal hidden context in Apple’s incomplete error data by automatically symbolicating unreadable symbols.
Handle iOS crashes with complete context using the React Native SDK while the Objective-C SDK runs in the background.
Fill in the blanks about iOS errors
See what the app was doing when the iOS error occurred: every user action, view controller changes, custom breadcrumbs.
Record events even when devices are offline or in airplane mode, then send errors as soon as connection is regained.
See the full picture of any Cocoa exception See the full picture of any iOS exception
Record environment and state details to recreate bugs down to the device, iOS version, app version, free memory, and disk space.
Sentry’s tag distribution graph also makes it easy to isolate and prioritize any iOS error by seeing how often it occurs in context.
- “How actionable is this error? Should I snooze the alert?”
- “Has an error with the same stack occurred before?”
- “Were they even close to catching the Aerodactyl?”
- “Was the device an iPhone 3GS or an iPad Air?”
- “Did they swipe left or swipe right?”
Resolve iOS iOS<< | https://sentry.io/for/cocoa/?code=objc | CC-MAIN-2018-17 | refinedweb | 270 | 57.67 |
I recently came across this JavaScript code:
function phonescrub(elmt){ str=elmt.value;str2="";ii=0; while (ii < str.length){ ch=str.charAt(ii); kk=0; while (kk < 10){ if (ch==""+kk){str2=str2+ch}; kk++; } ii++; } elmt.value=str2;}
function phonescrub(elmt){ str=elmt.value;str2="";ii=0; while (ii < str.length){ ch=str.charAt(ii); kk=0; while (kk < 10){ if (ch==""+kk){str2=str2+ch}; kk++; } ii++; } elmt.value=str2;}
It's a little difficult to decipher, so I'll interpret it.
Essentially, what this function does is it strips out all non-numeric characters.
This function has several issues (not the least of which is the variable names "str" and "str2"). But the main issue is the complexity. The method is 11 lines long and it took me 8 steps to explain what it is doing.
Using Regular Expressions, we can reduce 11 lines to 1:
function phonescrub(elmt){ elmt.value = elmt.value.replace(/[^0-9]/g, '');}
function phonescrub(elmt){ elmt.value = elmt.value.replace(/[^0-9]/g, '');}
Here is what this new function does:
Today I learned an important lesson about the ValidationSummary control. But first, some background.
The bug I was working on was this: the validation summary did not display the validation errors. The page would repost and redisplay, but there was no messaging as to what might be wrong.
The page was set up in a wizard sort of fashion with about three Panel controls that are displayed selectively, depending on which step the user is on. In order to prevent all the validation controls to fire every time a button is clicked, the various validators were assigned to different validation groups, one for each Panel.
It took me about an hour before I discovered this truth:
."
The first time I read it, I was in a hurry and didn't catch the very last part: "...that are not assigned to a validation group..." I guess it is important to read all the documentation.
My bug was that the ValidationSummary did not specify a validation group. According to the docs, that means that only validators that are not assigned a ValidationGroup will show up in the summary control. It was an easy fix (set the validation group when I set which panel to display), but it took me about an hour to figure it out.
I'm currently working on a project where we are storing quite a bit of data in Session for the user. Naturally, QA had a hard time testing whether or not something actually was stored in Session. So, we wrote a little page that showed the values of what we put in Session. The problem with this approach is that every time we added another item to Session, or removed an item, we would have to fix this page to include the new and remove the old. It got to be very tiresome.
I had just finished doing some work with reflection (my first time ever). And it occurred to me that I could use reflection to show everything that was stored in Session in such a way that it could be written once and never need to be updated.
Here's the basic rundown of what the utility does (see attached .zip file):
In the end, I always get to a native data type (or String or DateTime), so we wouldn't end up in an endless loop. Here's a sample of what it displays:
As an added bonus, it will also display the approximate size of all the objects (in bytes/kilobytes).
This utility has been very helpful for me, to make sure I know what is in Session, without having to step through the code in debug mode and stop and interrogate the Session object to find the specific value (or values) I'm looking for. This has also been very helpful to QA so that they can see what is put into Session.
At my day job, we're in the process of an enterprise re-write. The old stuff is a hodge-podge of Cold Fusion, Java, J# and C#. The website is Cold Fusion so the URLs often look something like this:
The new site is ASP.NET. Naturally, the URLs will look somewhat like this:
I'm sure that many of you have been involved in a scenario like this. The business wants to update the site. But we don't want to lose all the SEO ground we've gained over the years. So we have to support the legacy URLs.
There are different ways to tackle this problem: map the ".cfm" extension to be handled by ASP.NET using a custom IHttpHandler; or use an ISAPI filter; or use a third party URL Rewriter. In the end, most (if not all) of these methods rely on Regular Expressions.
Regular Expressions are powerful, but they can be confusing, and many developers avoid them. They're like the ultra-geeky kid in school. You don't really want to associate with him, but you'd love it if he was your lab partner, because you know he'd carry you through anything that you didn't understand.
The task to write the RegExes to match the Legacy URLs with the new URLs has fallen to me. And I'm actually happy about that. The best for me to learn anything is by doing it. So I get to sink my teeth into Regex and hopefully, I won't forget what I learned anytime soon.
And now, the good part - links:
I.
Utilities are among my favorites to code. Especially ones that are short and sweet and useful in many different situations. One of my favorites that I wrote was an Object Serializer.
Most examples of object serialization I have seen end up serializing an object to an XML file. That does have it's purpose. However, I've been in many situations where I'd like to serialize an object into XML and use the XML in code. Or I have XML in the form of a string which I need converted to an object.
The XmlSerializer is what we need. And if you look at the Serialize or Deserialize methods, you see that we have a lot of different options. We need some sort of Writer or Reader in order to do the conversion. Since I want to serialize to and deserialize from a string, the obvious choice was the StringWriter (to serialize) and the StringReader (to deserialize).
It is important to note that both the StringWriter and StringReader classes implement IDisposable and the documentation suggests that we should always dispose of our Reader/Writer when we're done with it. I'm a big fan of the "using" statement which automatically calls the Dispose() method when it leaves the using block.
I also decided to use generics so that I could have a strongly typed object going in and coming out.
So without further ado, here it is:
using System;
using System.Collections.Generic;
using System.IO;
using System.Text;
using System.Xml.Serialization;
namespace KevnRoberts.Utilities
{
/// <summary>
/// Used for serializing and deserializing objects
/// </summary>
/// <typeparam
name="T">The type of object to
de/serialize</typeparam>
public class Serializer<T>
{
/// <summary>
/// Converts the XML into an object
/// </summary>
/// <param
name="xml">The XML representation
of an object.</param>
/// <returns>An object of type T.</returns>
public static T Deserialize(string
xml)
{
//
convert to the serializable object
T item;
using
(StringReader reader = new StringReader(xml))
{
XmlSerializer
serializer = new XmlSerializer(typeof(T));
item =
(T)serializer.Deserialize(reader);
}
return
item;
}
/// <summary>
/// Converts an object into XML.
/// </summary>
/// <param
name="obj">The object to convert.</param>
/// <returns>The XML representation of the object.</returns>
public static string
Serialize(T obj)
{
StringBuilder
data = new StringBuilder();
using
(StringWriter writer = new StringWriter(data))
{
XmlSerializer
serializer = new XmlSerializer(typeof(T));
serializer.Serialize(writer,
obj);
}
return
data.ToString();
}
}
}
And here is an example of how you would use it:
string xml = @"<dateTime>2008-04-03T00:00:00</dateTime>";
DateTime deserializedDate = Serializer<DateTime>.Deserialize(xml);
DateTime date = new
DateTime(2008, 4, 3);
string serializedDate = Serializer<DateTime>.Serialize(date);
Of course, you'll want to use it to de/serialize something more complex than a date and time, but you get the point.
Dan Hounshell suggested that in the ASP.NET community, there are three types of people:
The Community Leaders, he writes, are the people who steer the community. He names Scott Guthrie as the primary steerer. The Practical Experts are people like himself. The ones that write books, present at events, have much read blogs.
Then there are the rest of us. He just calls us "Developers," but I like the term "Developer Barbarians." The Greeks called anyone who wasn't a Greek a Barbarian. Its the "there's Us and everyone who isn't Us" type thing.
Don't get me wrong. I'm not saying that anyone is condescending to us. And I'm certainly not saying that Dan is saying this. These are my words, not Dan's.
In fact, I like the fact that I'm a Developer Barbarian. I know enough to know that I don't know much. That is what makes me a barbarian. I aspire to move up to the "Practical Experts" level, but I'm not there yet.
I wouldn't say that I follow the community and don't study it. I do both. And I do actively seek out better ways to do things. But I'm not always in a position to follow it through to implementation. And most of the time, the "better way" that I found is something that came from the community.
One of the things that I love about software development is that I get to learn new stuff all the time. In fact, if I don't, I'll either search out new stuff to learn (if the job allows it) or I'll look for a new job. And, if time allows, I'll learn stuff on my own at home.
That's what I here to do. Learn. Because it's true. I'm a Developer Barbarian. But I don't want to stay that way. | http://weblogs.asp.net/kevnroberts/ | crawl-002 | refinedweb | 1,709 | 66.33 |
pun | v2 switch to v1
Version History
v2.40 (22. November 2021) Fixed: Unity 2021.2 support. PhotonEditorUtils.IsPrefab now uses the correct prefab stage utility for Unity 2021.2 and up. Changed: PhotonHandler.Dispatch is now catching exceptions. It will count them but only provide the first one's callstack (to avoid log spamming). Note: While the Photon lib caught exceptions in v4.1.6.10, that has not been a good solution. This new approach can be changed in source and is more flexible overall. Changed: On Xbox, the peer classes no longer assume a native socket plugin exists. This was setting the UDP implementation to null in some cases, which failed to connect. Updated: Library and Realtime to 4.1.6.11. More details in the change logs. v2.39 (21. October 2021) Changed: PhotonTransformView interpolation of position and rotation. It's now framerate independent. Updated: Library to 4.1.6.10. More details in changes-library.txt. v2.38 (12. October 2021) Updated: Library to 4.1.6.9. More details in changes-library.txt. v2.37 (27. September 2021) Added: CustomType deserialization length checks. Added: PhotonNetwork.EnableCloseConnection which defaults to false. This is a breaking change: Only if you set this to true, CloseConnection will work anymore. Updated: Library to 4.1.6.7. More details in changes-library.txt. v2.36 (20. September 2021) Updated: Realtime API. Check out changes-realtime.txt. Updated: Library to 4.1.6.6. More details in changes-library.txt. v2.35 (24. August 2021) Changed: CheckConnectSetupXboxOne() to allow AuthOnceWss and Datagram Encryption (GCM mode). It will set ExpectedProtocol and EncryptionMode silently. Datagram Encryption is only possible with a native plugin. Changed: Interest Culling in UtilityScripts/Culling/. Added: Rate limiting for interest group changes per second. No matter what the algorithm tells us, it makes sense to not change interest groups more than 4-6x per second. Changed: The GetActiveCells method to sort the "nearby" part of the cells list. This allows us to not change interest groups if just the order changed (due to "nearby" calculation, that could happen). Changed: The scene view grid now displaces the various area-indexes for cells that overlap (root, node and leaf). Note: Overall, the algorithm to select cells is just a quick showcase. It should avoid unsubscribing groups one by one and also avoid subscribing nearby groups one by one (while moving only along one axis of the areas). Fixed: A potential NullReferenceException for OwnershipUpdate events. If the PhotonView is unknown, this gets logged now. This is caused most frequently by loading scenes (or initializing them) without pausing the incoming message queue (dispatch). If ownership changed for objects, the Master Client will let new players know this but that failed, if the joining client didn't know the objects yet. Updated: Realtime API. Check out changes-realtime.txt. Updated: Library to 4.1.6.5. More details in changes-library.txt. v2.34.1 (28. July 2021) Updated: PhotonNetwork.PunVersion to "2.34.1" now. The previous version accidentally used "2.33.3" still. Updated: For Unity 2021, the obsoltete CompilationPipeline.assemblyCompilationStarted got replaced with CompilationPipeline.compilationStarted. v2.34 (26. July 2021) Changed: The PhotonView.Owner and .Controller for "room PhotonViews" will be changed properly whenever the Master Client switches. Room PhotonViews are identified by IsRoomView (which checks CreatorActorNr == 0). Changed: PUN needs to exit play mode in Editor, if there is a recompilation (due to code change, etc) to prevent running into endless NullReferenceExceptions and a flood of other errors. As the connection can't be held up after recompile, this is no big step back. The log will include a hint why the client exited playmode. Changed: The PhotonNetwork constructor creates a NetworkingClient (in Editor) to avoid exceptions due to possible recompilation. Changed: PhotonNetwork.StaticReset was running as RuntimeInitializeOnLoadMethod(RuntimeInitializeLoadType.BeforeSceneLoad) but will run earlier now, using AfterAssembliesLoaded. Updated: Library to 4.1.6.4. More details in changes-library.txt. v2.33.3 (07. July 2021) Fixed: Skipping cleanup of PhotonViews if they never had a legit ViewID. This could result in a NullReferenceException (e.g. when exiting play mode). Internal: RegisterPhotonView will set view.removedFromLocalViewList to false now. This is an internally used value. v2.33.2 (06. July 2021) Fixed: Assigning photonView.ViewID = 0 will remove the PhotonView from the list of active views. Added: Namespaces for various classes that were not having any. v2.33.1 (24. June 2021) Updated: Library to 4.1.6.3. More details in changes-library.txt. Note: This has no distinct value for PhotonNetwork.PunVersion and no separate release in the Asset Store (the previous release was on the same day). v2.33 (24. June 2021) Fixed: Calculation of PhotonView.AmOwner. It is possible that the owner is set to actorNumber 0, which instead will apply the Master Client as owner. This caused AmOwner to be wrong. Fixed: Controller updates when players enter a room. Every client has to update the views, as an owner may return and regain control. Fixed: ControllerActorNr assignment will check if the assigned controller is inactive. In that case, the Master Client will step in (right during assignment). Added: Additional check if the settings file exists, to avoid using CreateAsset, when it does. Added: JoinRandomOrCreateRoom(). This was implemented in Realtime but not PUN. Changed: The enumeration CustomAuthenticationType was changed to fix naming inconsistencies. Use PlayStation4 and PlayStation5 respectively. Old variants are available as obsolete. Fixed: In Asteroids Demo, remote ships are not reacting to input when remote controller is inactive. Fixed: Potential KeyNotFound issue in the Asteroids Demo, PlayerOverviewPanel if remote players left or the object got destroyed. Added: PlayerDetailsController.OnDisable() override (used in cockpit) to unsubscribe from callbacks which are subscribed in OnEnable(). Both override the inherited implementation and call base. Updated: Library to 4.1.6.2 with several smaller adjustments. More details in changes-library.txt. v2.32 (09. June 2021) Fixed: Editor scripts to not overwrite the PhotonServerSettings when upgrading projects or loading them without a library folder. While the Editor starts up, the settings may be blank but not null. As before, if the asset is not available, it gets created. Updated: Library to 4.1.6.1 with several smaller adjustments. More details in changes-library.txt. v2.31 (19. May 2021) Fixed: RpcList updating. Assumptions about editor-callbacks order were incorrect and caused a variable to not be set as needed. Changed: The PhotonViewHandler now resets the sceneViewId to 0 for prefabs. This avoids leaking sceneViewIds to the prefabs. Updated: Export platforms for .Net 3.5 lib. It was still enabled for some platforms accidentally. Now the minimum of data is in that meta file. Updated: Export platforms for .Net Standard 2.0 dll. It should be used in all platforms except .Net UWP. Internal: AccountService can now send an "origin" as reference to the package the registration originates from. Updated: Library to 4.1.6.0 with various changes for timeout detection and resends. v2.30 (13. April 2021) Changed: When the PhotonView.ControllerActorNumber should be set to 0, it now sets the backing field to the Master Client's actorNumber instead. This prevents misunderstandings when updates are sent with the Master Client's actual actorNumber Changed: Setting the PhotonView.ControllerActorNumber will now refresh the Controller in all cases. This makes sure that (e.g.) scene objects get a Controller value OnJoinedRoom (they might exist before a room is joined). Updated: EventSystemSpawner to log an error and pause the execution, if only the new input system is being used. The PUN demos are not compatible with it unless you set Active Input Handling to "Both". Note: Only the PUN demos are incompatible with the new input system. PUN itself is not affected. As the demos are in separate folders, you can easily delete all of them and use the new input system in your project. Updated: Realtime API to 4.1.5.4, preventing several error cases. v2.29 (15. March 2021) Updated: Minimum supported Unity version to 2018.4.x. We develop PUN 2 in Unity 2018.4.22 from now on. Also the language compatibility level '.Net Standard 2.0' is now being expected. Updated: Export settings for the dlls in Assets\PhotonLibs. All platforms will use .Net Standard 2.0, except the UWP export with .Net Runtime (which uses Metro with a placeholder being the lib in PhotonLibs\). Removed: Setting of protocol 1.6 when connecting directly to a Master Server. This was a workaround for older Photon Server SDKs. Note: If you host Photon yourself, you can set the LoadBalancingClient.SerializationProtocol to SerializationProtocol.GpBinaryV16, before you connect. Added: IPunOwnershipCallbacks.OnOwnershipTransferFailed. This helps react when an Ownership Transfer request failed (due to earlier changes). This might mean you have to implement OnOwnershipTransferFailed in some scripts. Changed: PhotonView OwnerActorNr and ControllerActorNr to make use of the new Room.GetPlayer() option (to get the Master Client instance if the owner/controller is set to 0). Fixed: The PhotonView's "Find" button was not saving the observed components in scenes, which meant the script wasn't called at runtime. Updated: Library to 4.1.5.2. v2.28.3 (03. March 2021) Fixed: PhotonView.AmOwner was not updated ever. v2.28.2 (03. March 2021) Fixed: A potential issue with the execution order of the PhotonView component. It now sets it's own execution order to -16000 to get initialized before other scripts (especially on scene loading). Changed: Conditionals and logging output for Xbox builds. In the Editor, it's impossible to get a token, so PUN should not enforce this. Changed: PhotonView.ControllerActorNr will make sure the PhotonNetwork.LocalPlayer is not null when accessing the actorNumber through that. Changed: Instantiate will set PhotonView lastOnSerializeDataSent and lastOnSerializeDataReceived to null (which keeps instantiation for reused objects cleaner). v2.28.1 (25. February 2021) Updated: Library to v4.1.5.1. which fixes a problem with IPv6 addresses which was introduced in v4.1.5.0. v2.28 (23. February 2021) Changed: PhotonView code to cache Owner and Controller, as well as storing the sceneViewId for objects which load with the scene. This is a bigger change and complex, so it may break some projects, despite the testing! Mail us, if you run into issues with ViewIDs, Ownership, etc: developer@photonengine.com. Added: DisconnectCause DnsExceptionOnConnect and ServerAddressInvalid (which might be renamed). Changed: Inspector to trim the values of AppId fields. Fixed: Likely fixes issue with Resources.Load failing to find prefabs in latest Unity versions. This is more of a workaround for Unity problems/changes. Changed: The re-usable list of PhotonViews in RemoveInstantiatedGO() is now cleared when it's no longer needed to avoid lingering PhotonView references (memory leaking). Added: OfflineMode to StaticReset(). So fast play mode (without domain reload) will not keep the client in offline mode. Changed: PhotonNetwork.Server for offline mode. It will now report the GameServer when an offlineroom is existing. Added: In OfflineMode, PhotonNetwork.LeaveRoom() will now also call the OnConnectedToMaster() callback. As the OfflineMode changes values immediately, the timing of callbacks might be slightly different than in online mode. Joining a room offline is instantaneous. Updated: Library to v4.1.5.0. which includes using the new Native Encryptor API (v2), a fix for NonAllocDictionary.Remove() and new StatusCode entries for DNS and IP issues. v2.27 (12. January 2021) Removed: "Simple" networking. Due to the progress on an upcoming networkin solution (Fusion), we will keep this set of components an external addon in maintenance mode (supported but not improved/extended). Updated: The Realtime API and Photon3Unity3d.dll to v4.1.4.9. v2.26.1 (15. December 2020) Note: This fix release does not have a new PhotonNetwork.PunVersion. The changes were just for the Editor. Fixed: Some edge cases for the included "Simple" addon. Added: Changelog for Simple. v2.26 (09. December 2020) Fixed: Possible IndexOutOfRangeException causing import issues when cached best region summary is not in expected format. Changed: Wizard will load Realtime AppId from settings, if present. This should avoid surplus "registrations". Changed: PhotonEditor will not call RegisterByEmail, if a request is pending some result. For that, the serviceClient is now a class variable and not a local one. Added: AccountService.RequestPendingResult. It's reset when the request ends (fail or success). Added: extra check to avoid sending useless (empty array payload) OwnershipUpdate event. Updated: The Realtime API and Photon3Unity3d.dll (v4.1.4.8) with smaller improvements and changes. Check the respective changelogs. v2.25 (30. November 2020) Added: Simple Network Sync for PUN2 a.k.a. "Simple" and PhotonUtilities. Note: extends PUN2 to operate on a simulation-based tick timing system that uses circular buffers. So far, Simple was distributed as addon, now it's included. Note: See: Updated: Realtime API in a few ways. See respective changelog. v2.24 (25. November 2020) Added: PhotonTransformView.UseLocal. It defaults to true for any new PhotonTransformView but allows to get the "old" behaviour with world position and rotation data being synced. Fixed: Inspector UI for PhotonTransformView and PhotonAnimatorView. Warning box for using triggers broke horizonal bounds, causing horizontal scrollbars on inspector. Switched to standard helpbox. Changed: Drastically lowered the MaxDatagrams value, to not send too much in one frame. Changed: Default sendFrequency to 33ms. PUN will send in more frames but less datagrams in each. Changed: SendRate does not affect the SerializationRate value anymore. Also, SerializationRate can be set independent from the SendRate. Updated: Documentation about SerializationRate and SendRate. This now explains more details about sending. v2.23 (17. November 2020) Changed: CustomTypes are now split into CustomTypes and CustomTypesUnity. The latter is part of the Realtime API and registers automatically for Unity projects (in LoadBalancingClient constructor). This only changes internals for PUN. The Player class is only a de/serializable type in PUN. Changed: DebugLog info in case an RPC was not found. Making it clear that RPCs should return null or an IEnumerator (if you enable RunRpcCoroutines). Changed: Creation of the locally used EventData instance. The Parameters are now readonly and non-null. Updated: Registration wizard to avoid multiple registration calls. Fixed: Callback registration updates when the owner changes (PhotonView.SetOwnerInternal()). Changed: PhotonServerSettings label for Appid of PUN. Was labelled as "Realtime" but there is a "PUN" type of appids which should be used. Changed: Various places in PUN to re-use boxed bytes (instead of causing boxing in those places). Added: Handing for compile defines PUN_DISPATCH_IN_FIXEDUPDATE and PUN_DISPATCH_IN_LATEUPDATE. It locks Dispatch calls to the respective monobehaviour methods (and timing). Use one or the other. Changed: Depending on PhotonNetwork.MinimalTimeScaleToDispatchInFixedUpdate and the Time.timeScale, PUN will dispatch incoming messages in FixedUpdate or in LateUpdate. Before it potentially was running in both MonoBehaviour callbacks. Changed: Description of PhotonNetwork.MinimalTimeScaleToDispatchInFixedUpdate to reflect and explain the change above. Changed: Loading and creation of the PhotonServerSettings file. We got several reports of errors with this but could not properly reproduce them. Several safeguards were added and in worst case, you might end up with two PhotonServerSettings files (if you had an old one). Added: BestRegion now checks if the operation GetRegions was successful. If it returned an error, it skips pinging (and the client will get disconnected anyways). Updated: The Realtime API and Photon3Unity3d.dll (v4.1.4.6) with lots of improvements and changes. Check the respective changelogs. v2.22 (07. September 2020) Changed: Default observe option for PhotonView is Unreliable On Change. Fixed: PhotonView.Get wouldn't work if object was disabled. This fixes that. Fixed: PhotonView.Get was using GetComponentInParent which will fail to find the PV if an object is disabled. Custom GetParentComponent solves that problem. Fixed: Code using the WebSocketSharp library. It's only for non-WebGL builds and needs to be put inside the conditional code of WebSocket.cs. Changed: Export settings for websocket-sharp.dll. It should export to any platform (if needed) except for console, web and Windows Store. Each such platform has a more specific solution. Added: AmOwner property to cached ownership values. Added: SetControllerInternal method for use with advanced SNS ownership handling. Not used currently by Pun2. Changed: Photonhander for client disconnects fixed to only change owner to null rather than Master. Changed: Only change owner of objects that will survive autoCleanup. Fixed: Mispelling of Owner as Onwer. Changed: Only rebuild controller cache on JoinRoom if client is the master. Otherwise leaves the controller as -1 indicating the current controller isn't known yet. Changed: PhotonView custom inspector. Default observe option for PhotonView is "Unreliable On Change" again. Added: OnSerializeView now checks components, so list items being null in Observables are not breaking anymore. Changed: A PhotonView now has IOnPhotonViewPreNetDestroy, IOnPhotonViewOwnerChange and IOnPhotonViewControllerChange callbacks that help with fine grained control of these cases. Interface IOnPhotonViewCallback is renamed to IPhotonViewCallback. Updated: The Realtime API and Photon3Unity3d.dll (v4.1.4.5) with lots of improvements and changes. Check the respective changelogs. v2.21 (19. August 2020) Fixed: Handling of case "blocked by Windows firewall". WebSocket-sharp does not call OnError in this case. This affects WebSocket.cs. Updated: AuthenticationValues.Token setter to be protected internal. Only needs to be set by Photon LoadBalancing API. Updated reference docs, too. Fixed: Supress a minor unused warning in SlotRacer. Fixed: When Player disconnects - only change the owner of that players created objects if AutoCleanup is false. Otherwise, do nothing as they are all about to get destroyed, and an OwnerChange to Master could break things for some users. Fixed: Duplicate call to StaticReset(). This also created the LoadBalancingClient two times and caused a callback registration to get lost. This fixes the reported problem with scene synchronization in 2019.4 and up. Changed: PhotonNetwork.LoadLevel() methods will skip loading, if PhotonHandler.AppQuits is true. This avoids a minor issue in the Editor when a script loads a scene in OnLeftRoom() while ending playmode. v2.20.2 (04. August 2020) Fixed: Destroy code that caused a "Ev Destroy Failed." error log. v2.20.1 (03. August 2020) Fixed: An issue with Standalone builds in Unity 2019.4.x, which did not connect. The main loop was somehow not called if PhotonMono was created in RuntimeInitializeOnLoad callback. Changed: RuntimeInitializeOnLoad() should only happen in the Editor, where we may disable the Domain Reload. Removed: Code that automatically added PV to gameobjects with MonoBehaviourPun (the PhotonView is not a must-have for those scripts (but typical)). Changed: PhotonTransformView is now sending local transform values. This should work better with parenting. v2.20 (03. August 2020) Fixed: Potential cause for NetworkingClient being null in Editor on enter playmode. The EditorApplication.isPlaying value has a race condition. Fixed: Edge case where ownershipCacheIsValid was not getting correctly updated for leaving players when Cleanup is disabled. Added: CustomContext for PhotonWizard AccountService creation. Added: Some additional Standard Assets for third person demos (For SNS overlap with existing demo). Note: ThirdPerson scripts namespace changed to avoid conflicts with Unity Standard Assets they are taken from. Added: "Chat Extended" features. Properties for public channels and users. Event OnErrorInfo. Wrapped in compile define. Added: Additional method for finding nested components. Used by SNS. Fixed: PhotonView.ResetOwnership will no longer run into NullReferenceException. Fixed: Another potential NullReferenceException due to bad placing of a null-check (in OnEvent, case OwnershipTransfer). Fixed: Asteroids scenes had missing (obsolete) component on cameras. Added: Auto-find Observable Components in PhotonViews and Multi-select. PhotonViews can now (optionally) find observable components on the same object, making the setup easier. Added: Initial support to turn off Domain Reloading, also known as Fast Play Mode. Please report issues by mail, should there be any. Changed: InstantiateSceneObject got renamed to InstantiateRoomObject, which is a better terminology for objects that belong to the room but are not loaded with the scene. Old naming is obsolete. Changed: Added argument to OnControllerChange callback - controllerHasChanged. Fixed: Support to network-destroy scene objects (loaded with the scene). Known limit: The events to remove networked scene-objects are fired on join, not on load. This may actually desync players, depending on when scenes get loaded. Added: Nullcheck to the StatesGui to avoid issues when recompiling during runtime (in Editor). Changed: Initialization of PhotonEditor to make sure it runs for the first time after the Editor finished loading. OnInitialHierarchyChanged registers for the events. Fixed: Potential issue when the background image could not be loaded and would assign to the style as null. Removed: PhotonEditor static constructor and InitializeOnLoad attribute (using InitializeOnLoadMethod exclusively). Removed: Support for PlaymodeStateChanged event as used before Unity 2017.2 (without the parameter). Affects PhotonEditor.cs. v2.19.3 (06. July 2020) Fixed: An issue with ownership, which was introduced in v2.19.2. v2.19.2 (02. July 2020) Fixed: Wasn't setting owner in NetworkInstantiate, which was causing a failure to reset old owner values on second room join. v2.19.1 (29. June 2020) Changed: Don't run GameObject removal code if app is quitting. Replaced: Mixamo animations with Standard Asset equivalents (for licensing reasons). Fixed: A problem in the Photon3Unity3d.dll (v4.1.4.3). New version: v4.1.4.4. Fixed: NonAllocDict. Capacity change was broken. This is a critical fix. PUN uses this internally for the PhotonViewCollection. Added: Indexer to NonAllocDict to allow Dict[key] get and set. Added: NonAllocDict.Clear() method. v2.19 (25. June 2020) Fixed: FindPunAssetFolder, which relied on a file that got renamed earlier. (no new asset file version) v2.19 (24. June 2020) Added: Documentation for PhotonView.Find(ViewID). Added: Callback "OnPreNetDestroy" and registration to PhotonView. If necessary, scripts can get a call before networked objects get destroyed. This may help unparent/rearrange objects. Added: Check that the Dev Region override is in the list of available regions (from Name Server). If not, the first available region is being used. Changed: RegionPinger.Start() now attempts to use a ThreadPool WorkerItem to ping individual regions. Only if that fails, a separate managed Thread is used. This should help avoid running into Thread limits (if any). Changed: The PhotonServerSettings field "Port" now gets used, even if the client connects to a Name Server. Before, the Port was only used if the checkbox "Use Name Server" was off. Added: Proxy setup for WebSocket connections. To actually use the proxy settings, a special WebSocket implementation is needed. In the PUN 2 package, this value won't do anything. Fixed: WebSocketTcp to call peerBase.OnConnect(), as expected by all IPhotonSockets. Changed: Error messages for RPC issues. Not found, parameters and too many components implementing them. Now uses GO as context (click the Editor log entry to highlight the exact target object). Changed: OnJoinedInstantiate component to add more features. It now checks Player.HasRejoined to decide if the local player needs to instantiate networked objects or if that was done before. Update: The link.xml to also cover the client lib namespace ExitGames.Client.Photon. If Unity fails to run a build with the error that some default constructor is missing, move the link.xml to your project root (or merge with yours). This avoids too much code stripping in IL2CPP builds. Fixed: AccountService Context description override to go with the authentication token for third party registrations. Added: TrafficRecorder with sample. See TrafficRecorderSrc for docs and usage. Fixed: NullReferenceException in PhotonTeamsManager when accessed in Awake. Added: EnableProtocolFallback in the PhotonServerSettigns. It is enabled by default, so the client will try another protocol if the initial connect to the Name Server fails. It's only applied in ConnectUsingSettings(). This is disabled for WebGL and Xbox One. Updated: The Realtime API and Photon3Unity3d.dll (v4.1.4.3) with lots of improvements and changes. Check the respective changelogs. Some important changes listed below: Improved: Performance in terms of GC. More internal objects are pooled. Events that contain just a byte[] can now be sent and received with almost zero allocations. Check LoadBalancingPeer.UseByteArraySlicePoolForEvents. Changed: Internals to avoid accumulation of Threads. Fixed: A rare threading issue for UDP connections. Fixed: Best Region selection for UWP builds on the .Net Runtime. Added: PhotonPeer.TrafficRecorder and ITrafficRecorder. This can be used to capture traffic on UDP connections. Added: Checks for connections on WebGL and Xbox One. Settings that can be corrected (to match the platform's requirements) are corrected with a log note. An Exception is thrown, if that's not possible. Connect methods may throw an Exception, which are to be fixed during development time. v2.18.1 (30. April 2020) Fixed: String serialization issue in the Photon3Unity3d.dll (v4.1.4.0). Updated to Photon3Unity3d.dll (v4.1.4.1). v2.18 (28. April 2020) Fixed: Support for Unity 2019.3. As Unity 2020 is still beta, there may be more import / build issues. Please report via mail. Note: We noticed upgrading projects may create more than one PhotonServerSettings file. If your AppId seems gone, look up these files and keep the one with settings. Fixed: Asteroid Demo bullet instantiation uses the position sent via the rpc, instead of the current local position. Updated: Nicer format for the countdown utility script to avoid floating points. Fixed: NullReferenceExceptions in PhotonTeamsManager.OnPlayerEnteredRoom and PhotonTeamsManager.OnPlayerLeftRoom. Added: PhotonTeamsManager.GetTeamMembersCount which is a helper method to get size of teams directly. Fixed: Photon teams extension methods for team assignment now rely only on Player.SetCustomProperties. Fixed: Removed obsolete usage of GuiLayer Components in demos for compatibility with Unity 2019.3 onwards. Fixed: Some warnings that Unity was showing on build. Changed: The checks before sending operations. As long as the peer is switching servers, there is no connection and operations can not be sent. This is logged better. Added: Logging of OnJoinRoomFailed to SupportLogger. Changed: StripKeysWithNullValues to reuse a list of key with null-values. This causes less allocations and garbage to clean. Fixed: Setting the serialization protocol for connections that directly go to the Master Server. This was broken in v2.17 by accident. Changed: The assembly PhotonWebsocket (WebSocket folder) will no longer export to UWP / Windows Store builds, even if the compile define WEBSOCKET is set. It is simply incompatible with the platform. Changed: LoadBalancingPeer.ConfigUnitySockets() will not set the UDP/TCP sockets for Windows Store builds. Changed: RegionHandler.GetPingImplementation will no longer use PingMono class, unless it is defined by some platform-specific code. Updated: The Realtime API and Photon3Unity3d.dll (v4.1.4.0) with lots of improvements and changes. Check the respective changelogs. v2.17 (26. March 2020) Added: New "Dev Region" feature. With this setting, all development builds will use the same region, avoiding initial matchmaking problems with best region selection. Development builds are turned on automatically, when the PhotonServerSettings get created initially and the "Dev Region" is set on first run in Editor. Changed: PhotonServerSettings inspector to accomodate the new values and reorganize existing ones. Changed: Instantiate and InstantiateSceneObject to not create a new object while not (yet) in a room. This avoids issues where an invalid showed up. There is a new debug log for this case, as it causes null-reference exceptions (in case it's used incorrectly). Added: New animations for the Basics Tutorial. The previous ones were causing warnings on import / reload. Added: New callback OnErrorInfo when the client receives an ErrorInfo event from the server. Implemented in MonoBehaviourPunCallbacks and SupportLogger. Changed: PhotonNetwork.SetPlayerCustomProperties return bool instead of void. Internal: ObjectsInOneUpdate will now initialize the capacity of a SerializeViewBatch. Removed: Metro folder and dlls. Unity deprecated builds for Windows 8.x Store, which means PUN can't support this anymore. Windows 10 UWP builds are supported using IL2CPP and the regular managed Photon dlls. Fixed: Unity Cloud Builds. A check prevents creation of ServerSettings when using Cloud Build. Internal: The PhotonEditor can now define a "customToken" for the AccountService registrations. Changed: Utility script OnJoinedInstantiate. It is now more versatile. Updated: SocketWebTcp.cs which should no longer define a SerializationProtocol. This is now part of the inherited class IPhotonSocket and automates sending the used protocol. Fixed: PunTurnManager ignores processed event callback when sender is -1, so that it can be instantiated even when player is not in room. Updated: Better warning for Trigger usages in PhotonAnimatorView inspector and at runtime. Updated: The Realtime API and Photon3Unity3d.dll (v4.1.3.0) with lots of improvements and changes. Check the respective changelogs. v2.16 (26. November 2019) Reverted: Some changes in the TypedLobby class. Setting Name and Type does not depend on some other state. Just make sure the Name is not null or empty or the TypedLobby will point to the "default lobby" with a fixed/defined behaviour. Fixed: Properly set the received interest group on the PhotonViews of the locally instantiated networked GameObject. Fixed: Issue in RegionHandler in WebGL which prevented connection to master server if a scene was loaded during best region ping/calculation. Fixed: A rare condition where a region preference can't be parsed into the region and ping values the inspector attempts to show. We check the result of the "split" now. Updated: The email check for registrations. Should be more accurate. Internal: Optimized Instantiation, RPC calling and sending with cached / reusable keys. This avoids some garbage creation. Added: Option to run RPCs as coroutines. Disable with PhotonNetwork.RunRpcCoroutines = false. This makes the upgrade from Pun Classic easier. Added: OpJoinRandomOrCreateRoom. When the random matchmaking does not find a suitable room, this operation creates one immediately. Added: The Scene Settings inspector now links to the online documentation for this topic. Click the "book" icon which Uniy uses to explain built in components. This works in other cases, too. Changed: LoadBalancingPeer and ChatPeer now look for SocketNativeSource instead of SocketWebTcpNativeDynamic when the target platform is XB One. A new Xbox addon is coming up on our SDK page. Updated: Photon3Unity3d.dll to v4.1.2.19. v2.15 (19. September 2019) Fixed: NullReferenceException when trying to log an error for a received RPC with a non existing method and zero arguments. Changed: Calling an RPC method with a null or empty method name will not be executed and error will be logged. Fixed: Callbacks deregistration and PlayersPerTeam clearing in PunTeams which caused KeyNotFoundException in some cases. Fixed: NullReferenceException in SupportLogger.TrackValues in Unity Editor after compilation by making sure only one instance of SupportLogger is in the scene and SupportLogger.TrackValues is properly stopped when the application stops or client disconnects. Changed: TypedLobby refactored to reflect the proper definition of the default lobby: a lobby that has a null or empty name, the provided type is ignored in this case and will be considered default type. Changed: RPC methods not found error logs explictily state that RPC methods search does not include static methods which are not supported as RPC methods. Updated: SceneSettings Inspector improved and SceneAsset field added to make reference to scenes easier. Updated: Code comments for documentation.: | https://doc.photonengine.com/zh-cn/pun/v2/reference/version-history | CC-MAIN-2022-05 | refinedweb | 5,131 | 52.87 |
Pololu Blog ».
Concept
During some of the previous line following contests several of the robots had a tendency to slide around the corners. There is an infamous F1 race car (the Brabham BT46B) that featured a large fan that extracted air from the underside of the car, creating a large downforce on the chassis which greatly increased the traction of the tires. The car was remarkably successful, though it had the side effect of exposing drivers to very high lateral acceleration. The design was quickly outlawed after a single race when competitors immediately started work on similar designs. Since there are no structurally weak primates to worry about in the cockpit of a line following robot, it seemed like it would be fun to create one that used the same principles. Thus, the idea for the Suckbot was born.
Chassis design
The chassis for the Suckbot is a wedge with centrally located fan that extracts air from a large cavity on the bottom of the robot. This bottom cavity covers several square inches of the surface it is sitting on, so even a fraction of a PSI drop in the static pressure in this pocket of air relative to the atmosphere outside it will create a large force normal to the surface (downforce). The front ball casters and rear drive wheels are mounted using slots to allow the air-gap between the robot and the surface it travels on to be tuned so that enough air can pass through to avoid stalling the fan without requiring a large volume of air movement to create the static pressure difference.
The line-sensors and a pair of ball casters are located at the front of the wedge and the drive wheels are located at the back. The electronics, battery, and drive motors are located between the drive wheels to move as much weight as possible close to the turning axis of the robot, which reduces its angular momentum and allows for quick turns.
The bottom of the robot is a continuous surface so that the pocket of low pressure air can be formed underneath, but the top sides of the wedge are trussed to provide a high strength for their weight. The bottom surface and truss beams are relatively thick (about 5 mm), but the 3D printing process used to manufacture the chassis creates a thin shell of solid material on the outside surfaces filled with a thin walled hexagonal honeycomb matrix on the inside, further saving weight.
The fan housing contains several ribs that primarily serve to hold the motor, but were included also in the hope that they would create a sort of vane-axial fan system to improve the static pressure developed by the fan.
The chassis was designed using SolidWorks and printed with ABS plastic on a modified FlashForge Creator 3D printer at SYN Shop (a local hackerspace). I don’t remember the exact print-time for the chassis, but it was something like 8 to 11 hours.
Mechanical systems
The suction fan uses a 30 mm, 8 blade rotor designed for ducted fans on RC airplanes. A brushless DC motor that is rated to run at 4500 rpm/volt is used to drive the fan, so it runs at about 30,000 RPM at the battery voltage. It would be really cool if the robot could also run on an inverted surface, but the Suckbot tips the scale at right about 300 g with the battery (75 g) and the largest load I’ve been able to lift with the suction is only about 140 g.
The two drive motors are Pololu high-power micro metal gearmotors with a 30:1 gear ratio. I decided on the gear ratio based on what was used on other robots in previous line following competitions. I designed a set of 70 mm main drive wheels to use some silicone wristbands as tires. The wheels are secured to the D-shaped output shafts on the motors using a captive nut and set screw. They were 3D printed using ABS plastic on the same printer as the chassis. I experimented with some Pololu 70×8mm wheels in my testing. The injection-molded Pololu wheels ran smoother, but the wider silicone wrist band tires had more traction and I liked having orange tires, so I stuck with them.
I also made and 3D printed some ball casters for the front, but when I turned on the fan suction, they didn’t roll very well and the robot moved much slower. I designed some mounts for a pair of Pololu 1/2″ plastic ball casters, which worked much better.
As I mentioned previously, the wheels and casters are mounted using slots that allow the ride height to be adjusted to fine tune the air gap under the robot. The final air gap used for the competition was about 1.3 mm, though I may try to lower this further. The mounts I designed to use the Pololu ball casters don’t allow me to go quite as low as my previous design.
Electrical systems
The microcontoller that runs the Suckbot is an ATmega328P using an Arduino Uno bootloader. There is a tutorial on the Arduino website that shows how to set up this chip on a solderless breadboard in a manner where it behaves like a standard Arduino Uno and Adafruit sells some handy stickers that can be placed directly on the AVR chip to indicate the Arduino pin mappings. The Suckbot initially used a 400-point solderless breadboard to hold most of the electronics, but a few days before the competition I switched over to a SmokingCircuits.com ProtoBoard (similar to Adafruit Perma-Proto Boards) that I got out of the vending machine at SYNShop.
To sense the line, the Suckbot uses an array of 6 Pololu QTR-1RC reflectance sensors mounted about 12 mm above the surface at the front of the robot. I made an add-on bracket for the chassis that allows the spacing between the sensors in the array to be adjusted. This bracket combined with the low ride height also shields the sensors from interference by ambient light.
The drive motors are controlled with a Pololu DRV8833 dual motor driver carrier. It is useful for fast line following to run the motor driver off a regulated power supply to ensure that the mechanical power produced by the motors from a given duty cycle doesn’t vary as the battery voltage changes, so a a Pololu adjustable step-up voltage regulator that is set to 8 V was used power the motor driver from the battery.
The fan motor is controlled by a Turnigy Plush 6A brushless speed controller. I had a lot of trouble getting this to work with my microcontroller and the servo signaling in the Arduino environment. I ended up using an N-Channel MOSFET as a switch across the ground of the ESC so that power to the ESC could be controlled by code on the microcontroller. I’ve detailed the problems I was having and included some demo code that works with my ESC at the end of this post in the hope that anyone else who is using an ESC like this and runs into a similar problem might find it useful.
The battery I used for the competition is a 1,300 mAh 2 cell series LiPo, though I have a smaller 500 mAh battery I used during some of the testing. There is a 5 V regulator on the control board that powers the microcontroller and sensors and the fan motor ESC is powered by the unregulated battery voltage.
Control algorithm
The Suckbot uses a simple PD loop (no I term) that controls the direction the robot is turning by controlling the power to the left and right drive motors. Feedback for the motors is provided from the QTR sensors at the front of the robot. The fan is turned on to a fixed speed a fraction of a second before the robot starts moving. I considered varying the fan power using the PD term since downforce is most helpful when the robot is changing its orientation, but I was concerned the throttle response wouldn’t be quick enough and that the torque from changing the fan speed would affect the robot’s stability.
The robot’s code contains parameters to set minimum and maximum drive motor speeds. The minimum speed I used in my final tuning is a reverse speed, which allows the Suckbot to make sharper turns. However, I had to be careful not to make the reverse speed too large; otherwise, the overall speed would be unnecessarily slowed when the robot oscillated on a line. In my final tuning, I used 800 as a minimum speed and 1400 as a maximum speed where 0 is full reverse, 1000 is stop, and 2000 is full forward.
Results
The Suckbot finished in the middle of the pack in the line following competition with a best 3-lap time of 31.8 seconds (about 0.93 m/s average speed). This wasn’t a bad showing for my first line following robot, but I was a bit disappointed because I ran into some problems in my code in the final tuning (integer overflows, I suspect) and had to use slower motor speeds during the contest to ensure the Suckbot would consistently finish the course. Even with the slower code, on one of the two line following courses in the competition there was a corner that consistently caused the robot to veer left off of the course when starting a right hand turn. This occurred despite otherwise following the course very smoothly. The robot used a low fan speed for the competition and I suspect it would have run just as well without it at those drive speeds.
Future plans
I feel like there is a lot of room for improvement in the Suckbot that can be realized by further tuning of the existing platform to take advantage of the additional traction from the fan suction. The existing code needs to be examined and modified to ensure there are no integer overflows. I recently started playing around with an ESP8266 WiFi-to-TTL Serial board, and I plan to use it with a Sharp digital distance sensor (and static position flags around the course) to get wireless telemetry data from the robot so I can tune it using my laptop as it runs the course.
Fan ESC initialization code
When I powered the Arduino and my Turnigy Plush 6A ESC at the same time and initialized the Arduino Servo library to control the ESC with the default timings (“Servo myservo;”) and then set the servo position to 0 (“myservo.write(0);”), the ESC beeped back an error code that indicated an invalid input signal. I experimented with sweeps to different servo positions in my code after the servo library was initialized, but that didn’t seem to work either. I wondered if the ESC was booting, looking for a signal, and throwing an error code while the Arduino was running through its initialization, so I added the N-Channel MOSFET as a switch across the ground of the ESC so I could switch it on and off in my software to be certain that the Arduino was sending a PWM signal to the ESC when it turned on. However, that alone did not fix the problem, and I also had to find a lower threshold for the initial servo position that the ESC recognized as valid.
I’ve included some code below which I used to successfully initialize and control my particular ESC using the Arduino environment and also includes some comments that indicate what the threshold values were for my particular ESC. I noticed in my research that there are versions of this controller that use slightly different microcontrollers than the one I got, so it seems like there are no guarantees that the firmware and the threshold values I found will be the same on your ESC. Hopefully this code will be helpful for you to use as a starting point to find your own threshold values.
#include <Servo.h> Servo myservo; const int pinFanPWM = 9; //I/O pin for the fan ESC servo signal const int pinFanSwitch = 4; //I/O pin for the MOSFET turning on the fan ESC void setup() { pinMode(pinFanPWM, OUTPUT); pinMode(pinFanSwitch, OUTPUT); digitalWrite(pinFanPWM, LOW); digitalWrite(pinFanSwitch, LOW); //Initialize the fan myservo.attach(pinFanPWM); myservo.write(10); //8 is the lowest threshold for the ESC to recognize a valid signal //during initialization. The top is around 65. delay(100); digitalWrite(pinFanSwitch,HIGH); //Turn fan ESC on via MOSFET delay(4100); //4000 or 4100ms for ESC to initialize myservo.write(74); //71 is the lowest threshold to make the fan turn on delay(500); myservo.write(10); //Stop the fan //Controller is now initialized. The fan can now be turned on delay(2000); myservo.write(100); //Turn fan on medium power delay(5000); myservo.write(180); //Turn fan on full power (The ESC recognized anything //above about 135 as full power) delay(5000); myservo.write(30); //Turn fan off again delay(2000); digitalWrite(pinFanSwitch,LOW); //Turn fan ESC off via MOSFET (This is not necessary and it will need //to be reinitiallized to use the fan again.) } void loop() { }
I used the larger battery for the contest because I was doing testing with it and adjusting parameters just before the contest started and I did not want to throw the robot off by changing to the smaller battery at the last minute. I did not extensively characterize the battery life, but the 500mAh battery was enough for at least a few 10's of laps.
There is a short clip of the robot running in the video for the competition at 2:33. Thanks for asking about it.
Nathan
I haven't published the files anywhere because they are kind of specific to the parts I used (like the fan blades and motor), there are some bugs in the design (like the front sensor mount being taped on) and in general, it doesn't seem like it is easy to modify STL files. Is there some easy way to work with STL files I'm unaware of?
-Nathan
I edit STL files with Tinkercad (). I'ts a very easy tool for creating and editing 3D models. It's capable of importing STL files to edit.
Maybe you already knew that tool, I kinda mastered it. :D
could you tell and could you explain more.
do you have any video.
i want to build this one.
Sorry, I guess I didn't explain that very well. I want to set up a Sharp distance sensor pointing off to the side of the robot and, when I'm tuning it, I can place objects to the side of the line following course so the sensor can detect them as it passes. That would allow lap times can be more precisely calculated to get a more accurate idea of how changing settings affects the lap time. I still haven't gotten around to doing this yet.
-Nathan
We noticed that you asked a similar question on a few of these line-follower blog posts. In order to keep the comments from getting too cluttered, we responded to your post here.
In the future, you might consider posting questions like these on our forum.
Brandon | https://www.pololu.com/blog/491/nathans-line-following-robot-suckbot | CC-MAIN-2022-05 | refinedweb | 2,586 | 63.73 |
More About the using Directive
As you know, the using keyword, followed by a suitable namespace name, allows us to abbreviate a class's long fully-qualified name when we refer to it in the source code. Consequently, if we were to use the classes from the two namespaces Sams.CSharpPrimerPlus.ElevatorSimulation and Sams.CSharpPrimerPlus.BankSimulation (defined in Listing 15.3) in the source code of Figure 15.1, we could use the short classnames of the classes belonging to these namespaces by including lines 1 and 2 shown in Figure 15.1.
Figure 15.1. The compiler cannot know to which namespace Building belongs.
But what if both namespaces referenced ...
Get C# Primer Plus now with O’Reilly online learning.
O’Reilly members experience live online training, plus books, videos, and digital content from 200+ publishers. | https://www.oreilly.com/library/view/c-primer-plus/0672321521/0672321521_ch15lev1sec6.html | CC-MAIN-2020-40 | refinedweb | 137 | 58.79 |
Writing a Simple Automated Test in FitNesse
Each column header in the fixture must map to a public method in BookRules. FitNesse concatenates words in the column header name, appropriately adjusting capitalization to meet Java camel case naming conventions. For example, the column header "daily fine" maps to the query method dailyFine().
The one line of code within dailyFine() interacts with code in the library system. Here, you're testing Java classes directly, but fixture code might also access a system through an interface layer such as a web service.
Now, try out the test. You click the Test button located in the left-hand margin. The results are shown in Figure 1. (If you see yellow, something is configured improperly. Double-check your path statements. Refer to the FitNesse online user's guide. Send me an email. Or post a message to the FitNesse Yahoo! group.)
Figure 1. FitNesse execution.
The green cells indicate that an expected cell value matched what the fixture code actually returned. The red cell beneath the column header checkout period? indicates an error—the test expected 20, but the system returned 21. You know the system is right in this case, so you edit the FitNesse page and change 20 to 21. You rerun the test by clicking the Test button. You see all green, and a summary showing "Assertions: 3 right, 0 wrong, 0 ignored, 0 exceptions."
Take a quick look at an additional fixture, one that can verify the late fine calculations. You'll create a test table to use a new fixture named fixtures.CheckinBook, You edit the test page, updating its contents to look like Listing 4.
Listing 4. An additional fixture.
!path c:Fitnessefitnesse.jar !path C:Documents and SettingsjlangrMy Documentsworkspace gamelanLibrarybin !path C:Documents and SettingsjlangrMy Documentsworkspace gamelanLibraryFixturesbin !|fixtures.BookRules| |daily fine?|grace period?|checkout period?| |10|3|21| !|fixtures.CheckinBook| |checkout date|due date?|checkin date|daysLate?|fine?| |12/1/2006|12/22/2006|12/22/2006|0|0| |12/1/2006|12/22/2006|12/23/2006|1|0| |12/1/2006|12/22/2006|12/25/2006|3|0| |12/1/2006|12/22/2006|12/26/2006|4|40| |12/1/2006|12/22/2006|12/27/2006|5|50|
The new table verifies the due date, days late, and fine amount for a book, given its checkout date and return date. Note that two of the column headers, checkout date and checkin date don't end with a ?. They're not queries; they instead represent input data elements. FitNesse uses data in these "setter" columns to initialize a public field in the fixture.
The fixtures.CheckinBook table in listing 4 contains five data rows that cover a number of data scenarios. When you click the Test button, FitNesse calls out to the fixture five times, once for each row in the table. For each row, FitNesse first initializes public fields in the corresponding fixture using each of the setter columns. For example, FitNesse sets the public field checkoutDate to the value 12/1/2006 for each row in the table. Once all setters execute, FitNesse executes each of the query methods.
The corresponding fixture code appears in Listing 5.
Listing 5. fixtures.CheckinBook.
package fixtures; import java.util.*; import com.langrsoft.app.*; import fit.*; public class CheckinBook extends ColumnFixture { public Date checkoutDate; public Date checkinDate; public Date dueDate() { Checkout checkout = new Checkout(new Book(), checkoutDate); checkout.returnOn(checkinDate); return checkout.getDueDate(); } public int daysLate() { Checkout checkout = new Checkout(new Book(), checkoutDate); checkout.returnOn(checkinDate); return checkout.daysLate(); } public int fine() { Checkout checkout = new Checkout(new Book(), checkoutDate); checkout.returnOn(checkinDate); return checkout.amountToFine(); } }
For each query method that executes, remember that FitNesse has already populated the public fields checkoutDate and checkinDate.
You run the test page to ensure everything passes.
As a programmer, you might balk at the design of the fixture. The fixture code creates a checkout object and calls its returnOn method three separate times, once for each of the three query methods. That's not very efficient!
The important point is that the test tables are an expression of the requirements of the system. They might not represent how you think users should interact with the system. Yet, that's how the customer who designed the test tables wants to think about things.
The test tables are your focal point for future negotiation. You can center debate and discussion around these tables. You might get the customer to agree to split the fixtures.CheckinBook table into two. Or, you might realize that you need to redesign the APIs into the system. Even better, going forward, you can consider that the FitNesse tests can help drive the development of the system.
You've just scratched the surface in terms of FitNesse capabilities. There are numerous fixture types, including a RowFixture type that allows you to verify collections of data. FitNesse also provides many ways to help make your tests more expressive. Read through the online users' guide and experiment with the FitNesse capabilities that it explains. You might also check out the book, Fit For Developing Software, by Rick Mugridge.
Download the Code
You can download the code that accompanies this article here.<< | https://www.developer.com/tech/article.php/10923_3649506_2/Writing-a-Simple-Automated-Test-in-FitNesse.htm | CC-MAIN-2018-26 | refinedweb | 872 | 59.19 |
A few years ago I looked into the possibility of calling functions from an external process, and turns out you can do that so I did some research and found some code which was later used in one of my old projects that did stuff with 32-bit processes. I decided to go back and look at that old project and see if I can do the same with 64-bit processes.
Well turns out that calling 64-bit functions externally isn’t any different than what you would do with 32-bit the only thing you need is for the external process to be running in 64-bit mode and of course use 64-bit instructions.
For our demonstration we will compile a 64-bit process that will have a function that is going to get called externally from our external C# function, so here is a simple program I wrote:
// ConsoleApplication2.cpp : This file contains the 'main' function. Program execution begins and ends there. // #include <iostream> bool wasCalled = false; void __fastcall my_function_to_call() { std::cout << "function called" << std::endl; wasCalled = true; } int main() { printf("my_function_to_call address: %p\n", my_function_to_call); // i love using printf and std::cout in the same function! while (!wasCalled) { } int number; std::cout << "please enter an interger value: "; std::cin >> number; std::cout << number << std::endl; }
The program will output the function address of my_function_to_call to save us some reverse-engineering headaches and the program will keep looping until wasCalled is set to true.
The address to our function is 0x7FF781D4139D, so we need to call 0x7FF781D4139D in order to get past the while loop in the program.
Here is how you'd call that function externally using C#:
First you will need to reference two important namespaces.
using System.Runtime.InteropServices; using System.Diagnostics;
We will use the DllImport attribute in System.Runtime.InteropServices to reference native functions that we will then use to call the function externally and System.Diagnostics to get the process we want to work with.
We will need three functions:
[DllImport("kernel32.dll")] internal static extern IntPtr VirtualAllocEx(IntPtr hProcess, IntPtr lpAddress, uint dwSize, int flAllocationType, int flProtect); [DllImport("kernel32.dll", SetLastError = true)] internal static extern bool WriteProcessMemory(IntPtr hProcess, IntPtr lpBaseAddress, byte[] lpBuffer, uint nSize, out UIntPtr lpNumberOfBytesWritten); [DllImport("kernel32.dll")] internal static extern IntPtr CreateRemoteThread(IntPtr hProcess, IntPtr lpThreadAttributes, uint dwStackSize, IntPtr lpStartAddress, IntPtr lpParameter, uint dwCreationFlags, out uint lpThreadId);
VirtualAllocEx allocates memory in the defined process, WriteProcessMemory writes to a specific memory address in the defined process and CreateRemoteThread creates a remote thread and that function will execute our assembly code.
Please read the following documentation pages provided by Microsoft as they explain it better than I ever will:
We will call the following op code:
byte[] opcode = new byte[] { 0x48, 0xB8, 0x9D, 0x13, 0xD4, 0x81, 0xF7, 0x7F, 0x00, 0x00, 0xFF, 0xD0, 0xC3 };
All of this hexadecimal mess can be translated to:
48 B8 9D 13 D4 81 F7 7F 00 00 mov rax, 0x7FF781D4139D ; move the address 0x7FF781D4139D to rax FF D0 call rax ; call the value stored in rax registry C3 retn ; returns, without this opcode the process will crash
And when we put it all together we get this:
static void Main(string[] args) { IntPtr addr = IntPtr.Zero; UIntPtr bytesWritten = UIntPtr.Zero; IntPtr hThread = IntPtr.Zero; Process process = Process.GetProcessesByName("ConsoleApplication2")[0]; // the process we want to inject our opcode to byte[] opcode = new byte[] { 0x48, 0xB8, 0x9D, 0x13, 0xD4, 0x81, 0xF7, 0x7F, 0x00, 0x00, // mov rax, 0x7FF781D4139D 0xFF, 0xD0, // call rax 0xC3 // retn }; // allocate memory if((addr = VirtualAllocEx(process.Handle, IntPtr.Zero, (uint)opcode.Length, 0x00001000, 0x0040)) == IntPtr.Zero) { Console.WriteLine("failed to allocate memory."); return; } // write memory if(!WriteProcessMemory(process.Handle, addr, opcode, (uint)opcode.Length, out bytesWritten)) { Console.WriteLine("could not write to process' memory."); return; } uint lpThreadId = 0; // create thread if((hThread = CreateRemoteThread(process.Handle, IntPtr.Zero, 0, addr, IntPtr.Zero, 0, out lpThreadId)) == IntPtr.Zero) { Console.WriteLine("hThread value is 0x0."); return; } }
That's basically it though I haven't covered what to do with different convention calls, closing handles externally and this beautiful thing called stack alignment but this is your problem now.
See you next time when I figure out how to call game functions in an emulator. | https://dennisstanistan.com/blog/291/how-to-call-64-bit-functions-externally-using-c/ | CC-MAIN-2020-40 | refinedweb | 711 | 52.8 |
import * as Y from 'yjs'const ydoc = new Y.Doc()// You can define a Y.Map as a top-level type or a nested type// Method 1: Define a top-level typeconst ymap = ydoc.getMap('my map type')// Method 2: Define Y.Map that can be included into the Yjs documentconst ymapNested = new Y.Map()// Nested types can be included as content into any other shared typeymap.set('my nested map', ymapNested)// Common methodsymap.set('prop-name', 'value') // value can be anything json-encodableymap.get('prop-name') // => 'value'ymap.delete('prop-name')
ymap.doc: Y.Doc | null (readonly)
The Yjs document that this type is bound to. Is
null when it is not bound yet.
ymap.parent: Y.AbstractType | null
The parent that holds this type. Is
null if this
ymap is a top-level type.
ymap.set(key: string, value: object|boolean|string|number|Uint8Array|Y.AbstractType)
Add or update an entry with a specified key. This method works similarly to the Map.set method. The value can be a shared type, an Uint8Array, or anything JSON-encodable.
ymap.get(key: string): object|boolean|Array|string|number|Uint8Array|Y.AbstractType
Returns an entry with the specified key. This method works similarly to the Map.get method.
ymap.delete(key: string)
Deletes an entry with the specified key. This method works similarly to the Map.delete method.
ymap.has(key: string): boolean
Returns true if an entry with the specified key exists. This method works similarly to the Map.has method.
ymap.toJSON(): Object<string,object|boolean|Array|string|number|Uint8Array>
Copies the
[key,value] pairs of this Y.Map to a new Object. It transforms all shared types to JSON using their
toJSON method.
ymap.forEach(value: any, key: string, map: Y.Map)
Execute the provided function once on every key-value pair.
ymap[Symbol.Iterator]: Iterator
Returns an Iterator of
[key, value] pairs. This allows you to iterate over the
ymap using a for..of loop:
for (const [key, value] of ymap) { .. }
ymap.entries(): Iterator
Returns an Iterator of
[key, value] pairs.
ymap.values(): Iterator
Returns an Iterator of values only. This allows you to iterate through the values only
for (const value of ymap.values()) { ... } or insert all values into an array
Array.from(ymap.values()).
ymap.keys(): Iterator
Returns an Iterator of keys only. This allows you to iterate through the keys only
for (const key of ymap.keys()) { ... } or insert all keys into an array
Array.from(ymap.keys()).
ymap.clone(): Y.Map
Clone all values into a fresh Y.Map instance. The returned type can be included into the Yjs document.
ymap.observe(function(YMapEvent, Transaction))
Registers a change observer that will be called synchronously every time this shared type is modified. In the case this type is modified in the observer call, the event listener will be called again after the current event listener returns.
ymap.unobserve(function)
Unregisters a change observer that has been registered with
ymap.observe.
ymapmap.unobserveDeep(function)
Unregisters a change observer that has been registered with
ymap.observeDeep.
ymap.observe(ymapEvent => {ymapEvent.target === ymap // => true// Find out what changed:// Option 1: A set of keys that changedymapEvent.keysChanged // => Set<strings>// Option 2: Compute the differencesymapEvent.changes.keys // => Map<string, { action: 'add'|'update'|'delete', oldValue: any}>// sample code.ymapEvent".
ymapEvent.keysChanged: Set<string>
A Set containing all keys that were modified during a transaction.
See Y.Event API. The rest of the API is inherited from Y.Event. | https://docs.yjs.dev/api/shared-types/y.map | CC-MAIN-2021-39 | refinedweb | 579 | 53.27 |
In this article, we will discuss Matplotlib Arrow() in Python. Matplotlib is a powerful plotting library used for working with Python and NumPy. And for making statistical interference, it is necessary to visualize data, and Matplotlib is very useful. Furthermore, it provides a MATLAB-like interface only difference is that it uses Python and is open source.
Matplotlib Arrow function creates an arrow from one point to another point in a graph. This arrow can be useful to pinpoint graph items and increase the engagement of the graph. Moreover, every arrow has its unique properties to customize it. In this post, we’ll look at the arrow function in detail.
Syntax of Matplotlib Arrow() in python:
matplotlib.pyplot.arrow(x, y, dx, dy, **kwargs)
Parameters:
- x and y are the coordinates of the arrow base.
- dx and dy are the length of the arrow along the x and y-direction, respectively.
- **kwargs are optional arguments that help control the arrow’s construction and properties, like adding color to the arrow, changing the width of the arrow, etc.
- Constructor Arguments for **kwargs: width: float (default: 0.001) – width of full arrow tail
- length_includes_head: bool (default: False) – True if head is to be counted in calculating the length.
- head_width: float or None (default: 3*width) – Total width of the full arrow head
- head_length: float or None (default: 1.5 * head_width) – Length of arrow head
- shape: [‘full’, ‘left’, ‘right’] (default: ‘full’) – Draw the left-half, right-half, or full arrow
- overhang: float (default: 0) – Fraction that the arrow overhangs (0 overhang means triangular shape). Can be negative or greater than one.
- head_starts_at_zero: bool (default: False) – If True, the head starts at coordinate 0 instead of ending at coordinate 0.
Return of Matplotlib Arrow() function:
Returns an Arrow depending on our desired parameters that we provide as inputs.
Plot Arrow using Matplotlib Arrow() in Python
import matplotlib.pyplot as plt #define two arrays for plotting A = [3, 5, 5, 6, 7, 8] B = [12, 14, 17, 20, 22, 27] #add arrow to plot plt.arrow(x=4, y=18, dx=2, dy=5, width=.06) #display plot plt.show()
Output:
Explanation:
Firstly, an arrow is created where x and y parameters are set to 4 and 8. Setting dx and dy to 2 and 5 creates the length of the arrow. In addition, the width is passed as **kwargs constructor argument, which controls the arrow’s width to be created.
Note: dx is 0 to create a vertical arrow and dy to 0 to create a horizontal arrow.
Arrow styling using Matplotlib Arrow()
import matplotlib.pyplot as plt #define two arrays for plotting A = [3, 5, 5, 6, 7, 8] B = [12, 14, 17, 20, 22, 27] #add arrow to plot plt.arrow(x=4, y=18, dx=0, dy=5, width=.08, facecolor='red') #display plot plt.show()
Output:
Explanation:
Surprisingly, there are various ways of styling arrows. These different styling properties are applied by passing it as **kwargs argument. To demonstrate, the face color of the arrow being created is set to red. Here we get a vertical arrow as we have set the value of dx to 0.
Adding Annotations to Arrows
import matplotlib.pyplot as plt #define two arrays for plotting A = [3, 5, 5, 6, 7, 8] B = [12, 14, 17, 20, 22, 27] #add arrow to plot plt.arrow(x=4, y=18, dx=0, dy=5, width=.08) #add annotation plt.annotate('General direction', xy = (3.3, 17)) #display plot plt.show()
Output:
Explanation:
Annotation, in the simplest form, means adding a text at a point (x, y). To annotate an arrow means adding text around the arrow at a specific x and y coordinate value.
Syntax of Annotate function:
matplotlib.pyplot.annotate(text, xy,*args,**kwargs)
Where text to be added x and y are the point to annotate and, *args and **kwargs are optional parameters that control annotation properties. To summarize, ‘General direction’ text is added at x = 3.3 and y = 17.
Conclusion
To conclude, this article provides an obvious understanding of Matplotlib arrow() in python. All are included in this article, from plotting an arrow to styling it and adding annotations. Implementation of arrows in the python program along with examples is also well defined here.
However, if you have any doubts or questions do let me know in the comment section below. I will try to help you as soon as possible.
Happy Pythoning! | https://www.pythonpool.com/matplotlib-arrow/ | CC-MAIN-2021-43 | refinedweb | 742 | 66.44 |
Find Devices by OUI Connected to Local Network using Nmap and Python
A MAC (Media Access Control) address is a unique number assigned to a network adapter. In most cases, MAC address is displayed as a string of six hexadecimal numbers which separated by colons (e.g.
DC:A6:32:E1:5B:4B). First three hexadecimal numbers (e.g.
DC:A6:32) is called OUI (Organizationally Unique Identifier). OUI identifies the manufacturer.
This tutorial provides example to find devices by OUI connected to local network using Nmap and Python.
Prepare environment
Make sure you have installed Nmap and Python on your system. On Windows, make sure that Nmap directory is added to the PATH environment variable.
Using
pip package manager install
python-nmap library that allows to use Nmap and get scan results:
pip install python-nmap
Code
We have defined
target variable that specifies IP range of the network from 192.168.0.0 to 192.168.0.255. To determine that, use
ip command on Linux or
ipconfig command on Windows.
In our case, we will try to find Raspberry Pi boards connected to local network. OUI of Raspberry Pi manufacturer is
B8:27:EB or
DC:A6:32.
Nmap accepts
-sn option. It means that Nmap will scan devices on the network without port scan phase. On Linux, Nmap command must be executed with sudo to include the MAC address of each device in scan results.
import nmap target = '192.168.0.0/24' oui_list = ['B8:27:EB', 'DC:A6:32'] scanner = nmap.PortScanner() scanner.scan(target, arguments='-sn', sudo=True) hosts = [] for host in scanner.all_hosts(): addresses = scanner[host]['addresses'] if 'mac' not in addresses: continue oui = addresses['mac'][:8] if oui in oui_list: hosts.append(addresses) print(hosts)
If devices has been found, you will get results in the following form:
[{'ipv4': '192.168.0.100', 'mac': 'DC:A6:32:E1:5B:4B'}, {'ipv4': '192.168.0.123', 'mac': 'B8:27:EB:65:A1:2C'}] | https://lindevs.com/find-devices-by-oui-connected-to-local-network-using-nmap-and-python/ | CC-MAIN-2021-31 | refinedweb | 330 | 59.5 |
// Read words and print them in reverse order. // Variation 1: Fixed array sizes, use new to copy word. // Fred Swartz 2001-11-08, 2001-12-04 #include <iostream> // for cin, cout #include <cstring> // for strlen, strcpy using namespace std; int main() { char *allwords[1000]; // array of POINTERS to char strings char word[500]; // input buffer for longest possible word. int n = 0; // count of number of words. // read words/tokens from input stream while (cin >> word) { allwords[n] = new char[strlen(word)+1]; // allocate space strcpy(allwords[n], word); // copy word to new space n++; } cout << "Number of words = " << n << endl; // write out all the words in reverse order. // The dynamically allocated space is freed after the word // is printed, and the pointer is set to NULL. This isn't // necessary here because the program immediately terminates, // but it's a good, safe practice. for (int i=n-1; i>=0; i--) { cout << allwords[i] << endl; // print the word delete [] allwords[i]; // free space allwords[i] = NULL; // remove pointer } return 0; }//end mainThe big problem with this program is that is is subject to buffer overflow bugs -- words larger than 499 characters or more than 1000 words will simply overflow the arrays until something so serious happens that the program can't continue running. Let's hope it didn't overwrite one of you open output file buffers, for example. | http://www.fredosaurus.com/notes-cpp/datastructs/ex1/usingarray.html | CC-MAIN-2013-20 | refinedweb | 230 | 65.15 |
Chapter 16
Chapter 16
Using JavaScript and Forms
JavaScript wears many hats. You can use it to create special effects. You can use it to make your HTML pages "smarter" by using its decision-making capabilities. And you can use it to enhance HTML forms. This last application is of particular important and merit. Of all the hats JavaScript can wear, its form processing features are among the most sought and used.
Without JavaScript, forms are pretty much the domain of CGI programs and scripts. You use HTML to create the form, but the form itself is accepted and processed by some program running on the computer. With JavaScript you can process simple forms without invoking the server. And when submitting the form to a CGI program is necessary, you can have JavaScript take care of all the preliminary requirements, like validating input to ensure that the user has dotted every i.
Using JavaScript with forms is of such important that there are two chapters on this topic: the one
you're reading now, and Chapter 18, "Using CGI with JavaScript." This chapter details how to
interface a form created with standard HTML tags with JavaScript, and how to perform tasks
such as input validation. Chapter 18 is devoted to the CGI side of things, linking a JavaScript-enhanced form to a CGI program or script that runs on the server.
Creating the Form
There are few differences between a straight HTML form and a JavaScript-enhanced form. The main difference is that a JavaScript form relies on one or more "on..." event handlers. These invoke a JavaScript action when the user does something in the form, like clicks a button. The "on..." event handlers, which are placed with the rest of the attributes in the <FORM> tags, are invisible to a browser that don't support JavaScript. Because of this you can often use one form for both JavaScript and non-JavaScript browsers (though I don't necessarily recommend this on a support standpoint; you are usually better off creating two different forms).
Typical form objects include the following (I call them objects rather than the oft-used "widgets," because that's how JavaScript treats them):
- Text box for entering a line of text
- Push button for selecting an action
- Radio buttons for making one selection among a group of options
- Check boxes for selecting or deselecting a single, independent option
I won't bother enumerating all the attributes of these elements, and how to use them in HTML. Most any reference on HTML will provide you with the details (if you need a review, I include a summary of most all HTML 2.0 and Netscape-supported HTML 3.0 tags in Chapter 22, "All About HTML").
For use with JavaScript, you should always remember to provide a name for the form itself, and each element (object) you use. The names allow you to reference the object in your JavaScript program.
The typical form looks like this. Notice I've provided NAME= attributes for all form objects, including the form itself (which is also treated as a JavaScript object):
form is not designed to submit anything, the typical URL for the CGI program is omitted.
- METHOD="GET" defines the method data is passed to the server when the form is submitted. As this form is not going to submit anything, in this case the attribute is actually wasted puffery. It's shown here for example purposes.
- INPUT TYPE="text" defines the text box object. This is standard HTML markup here.
- INPUT TYPE="button" defines the button object. This is standard HTML markup except for the onClick handler.
- onClick="testResults(this.form)" is called an event handler -- it handles an event, in this case clicking the button. When the button is clicked, JavaScript executes the expression within the quotes. The expression says to call the testResults function elsewhere in the script, and pass to it the current form object (accomplished with the this.form parameter).
Getting a Value from a Form Object
Here's the full script you can try as you experiment obtaining values from form objects. Load the page, then type something into the text box. Click the button, and what you typed is shown in the alert box. The output of this script is shown in Figure 16-1.
testform.html <CD> >
<Figure 16.1. The form test input script.>
Here's how it works: You already know what happens when you click the button, from the paragraphs above: JavaScript calls the testResults function. The testResults function is passed the form object. I've given this object the name form inside the testResult function, but you can name this object anything you like.
This bears repeating in slightly different terms: the onClick="testResults(this.form)" passes the current form to the testResults function. In the testResults function this object goes by the name form. You do not use "this.form" within the testResults function, because JavaScript won't know what the "this" belongs to.
The function is simple -- it merely copies the contents of the text box to a variable named TestVar. Notice how the text box contents was referenced. I defined the form object I wanted to use (called form), the object within the form that I wanted to use (called inputbox), and the property of that object I wanted (the value property).
var TestVar = form.inputbox.value;Strictly speaking you don't have to pass the form object to a function in order to use that object within the function. Because the form is named you can reference it explicitly as a member of the document object. The following expression does the same thing as the one above: var TestVar = document.myform.inputbox.value; Here's how it breaks down, going from right to left: value is a property of the inputbox form control. inputbox is a member of the myform form object myform is a member of the current document.
Setting a Value in a Form Object
The value property of the inputbox, shown in the above example, is both readable and writable. That is, you can read whatever somebody types into the box, and you can write data back into it. The process of setting the value in a form object is just the reverse of reading it. Here's a short example to demonstrate. It's pretty much the same as the previous example, except this time there are two buttons, and the buttons have been relabeled. Click the "Read" button and the script reads what you typed into the text box. Click the "Write" button and the script writes a particularly lurid phrase into the text box.
set_formval.html <CD> (yuk) "Have a nice day!" in the text box.
Reading Other Form Object Values
The text box is perhaps the most common form object you'll read (or write) using JavaScript. However, you can use JavaScript to read and write values these form objects as well :
-.
For server/client communications, the server can send data to the client, storing a special value in a hidden field that the user doesn't see. For example, the value might be the number of times the user has submitted a form in the same session. If it's more than say, five times, the server knows not to accept any more entries from that user. The submission count is stored in a hidden field.
Hidden fields are particularly handy for storing temporary data that your JavaScript program may
need. Store the data in a hidden field, and it stays as long as the document remains loaded.
(However, note that the contents of hidden fields are lost when a document or frame is reloaded
or resized.) of buttons out of them;>The onClick event handler in this markup acts as a work-around for a bug in Netscape 2.0. Without the onClick event handler JavaScript will return the buttons in reverse order: 3, 2, 1, 0 -- instead of the proper 0, 1, 2, 3. The onClick handler does nothing, as its value is set to zero. <CD> >If you want to have more than one set of radio buttons in your form, just give them different names. For example, the first group of buttons might be named "rad1," and the second "rad2." The browser will allow you to select a button from each group.
Setting a radio button selection is even easier. If you want the form to initially appear with a given radio button selected just added the CHECKED attribute to the HTML markup for that button:
<INPUT TYPE="radio" NAME="rad" Value="rad_button1" CHECKED onClick=0><BR>
You can also set the button selection programmatically with JavaScript, using the checked property (there is also a selected property, but it is not properly functioning in all platforms supported by Netscape 2.0). Just specify the index of the radio button array you want to select. or not using the checked property. Likewise, you can set the checked property to add or remove the checkmark from a check box.
form_check.html <CD> > </HTML>
As with the radio button object, add a CHECKED attribute to the HTML markup for that check box if you wish to set a check box when the form first appears.
<INPUT TYPE="checkbox" NAME="check1" Value="0" CHECKED>Checkbox 1<BR>
You can also set the button selection programmatically with JavaScript, using the checked property. Just specify the name of the checkbox you want to check. Remember that you can check multiple check boxes.. Example:
form_textarea.html <CD> . You can enclose the text between the <TEXTAREA> and </TEXTAREA> tags, as shown in Figure 16-2. This method is useful if you wish to include hard returns, as these are retained in the text area box. Or, you can set it programmatically with JavaScript (but in this case you cannot add hard returns). Here's an example of the first method:
<TEXTAREA NAME="myarea" COLS="40" ROWS="7"> Initial text displayed here </TEXTAREA>
Using JavaScript you set the text area value:
formname.textarea.value = "Text goes here";
- formname is the name of the form.
- textarea is the name of the textarea.
- "Text goes here" is the text you want to display
<Figure 16-2. Default text can be placed between <TEXTAREA> tags.>
Text you write to the text area is treated as "unformatted." Any markup tags you use -- such as <P> or <BR> for line breaks, appears as-is in the text area box.
Using Selection Lists
List boxes let you pick the item you want out of a multiple-choice box. The listbox itself is created with the <SELECT> tag, and the items inside it are created by one or more <OPTION>..
Example of drop down list (displays only one item unless you click on the list):
<SELECT NAME="list"> <OPTION>This is item 1 <OPTION>This is item 2 <OPTION>This is item 3 <OPTION>This is item 4 </SELECT>
Example of drop down list (displays all four items):
<SELECT NAME="list" SIZE="4"> <OPTION>This is item 1 <OPTION>This is item 2 <OPTION>This is item 3 <OPTION>This is item 4 </SELECT>
Use the selectedIndex property to test which option item is selected in the list. The item is returned as an index value, with 0 being the first option, 1 being the second, and so forth (if no item is selected the value is -1). Here's a working example:
form_select.html
<CD> ); }
Other Things You Can Do With <SELECT> Objects
The <SELECT> object is the er, object of a lot of confusion to many JavaScript programmers. So here's a rundown of what you can do with the list boxes created with the <SELECT> object. In all of the below, formname is the unique name of the form, and selectname is the unique name of the select object you want to test.
- Determine the index of the selected option with:
index is the of the option that is selected (numbering starts at 0)
- Set the selected option with:
index is the index value of the option you want selected (numbering starts at 0)
- Determine the text of the selected item with:
Result = formname.selectname.options[Item].text;
- Determine how many items are in the <SELECT> object with:
- Set the text of a select option, as long as you also tell JavaScript to reload the page:
history.go(0);
index is the index of the option that you want to change
"Text" is the text you want to change to
Testing for Multiple Selections With the selected Property
You can allow for multiple selections in a list by using the MULTIPLE attribute in the <SELECT> tag. The JavaScript to process this is a bit more involved, because you need to test each item of the list to see if it's selected. A perfect way to do this is to use a for counter that enumerates through all the <OPTION> tags in a <SELECT> object (you control the number of iterations of the for loop by using the length property of the select object).
For example, say you have a list with five items on it, and the user selects the first and fourth item. The following JavaScript program will display:
Option 0 is selected
Option 3 is selected
in the alert box. Once again note that the option items are numbered starting from zero. The first option is 0, the second is 1, and so forth.
form_multiple.html <CD> <HTML> <HEAD> <TITLE>Multiple Selection Test </TITLE> <SCRIPT LANGUAGE="JavaScript"> function selectedItem (form) { var <INPUT TYPE="button" NAME="button3" Value="Test" onClick="selectedItem(this.form)"> <SELECT NAME="list" SIZE="5" MULTIPLE> <OPTION>This is item 1 <OPTION>This is item 2 <OPTION>This is item 3 <OPTION>This is item 4 <OPTION>This is item 5 </SELECT> </FORM> </BODY> </HTML>
Other Events You can Trigger Within a Form
I've used the onClick event handler in all of the examples in this chapter because that's the one you are most likely to deal with in your forms. Yet JavaScript supports a number of other event handlers as well. Use these as the need arises, and the mood fits. chapter).
Submitting the Form to the Server
In all of the examples above I've limited the action of the form to within JavaScript only. Many forms are designed to send data back to the server. This is called : it should call the mailMe() function, where the fields are appended to a mailto: URL. Netscape automatically opens a new mail window with the fields filled in. Write the body of the message, and send the mail off to the recipient.
onsubmit.html <CD> <CD> <HTML> <HEAD> <TITLE>
I'll reserve the remaining full details of using the onSubmit and submit instructions for Chapter 18, "Using CGI with JavaScript." At first blush it looks like you can use the mailto: URL as the ACTION string for the form. This should have the effect of automatically sending the contents of the text boxes straight to the recipient's mail box. The syntax looks like this:
<FORM NAME="testform" ACTION="mailto:someone@domain.com">In fact, this method does work with the initial release of Netscape 2.0. But since it is possible to submit a form without the user knowing (using the submit method), Netscape has removed this functionality in subsequent releases. This is a minor security issue where e-mail addresses can be picked up and mailed back to someone. These e-mail addresses could then be used for advertising or marketing purposes.
Validating Form Data Using JavaScript.
As you've read earlier in the chapter, forms on the Web consist.
Input Validation Routines
Most form validation chores revolve around basic data checking: does the user remember to fill in an box? Is it have the right length? Does it contain valid characters? With most forms you can readily answer these questions with a small handful of validation routines. You can write them yourself, or use the over two dozen validation routines included with this book.
A typical validation routine is determining if an input box contains only numeric digits.. Here's an example:
valid_simple.html <CD> >
Note that the isNumberString function is also smart enough to check for an empty input box. It
treats an empty box as an invalid entry, returning a 0. If this test were not provided a blank entry
would be treated as valid input.
Overview of Validation Routines
Here is a run down of some of the other input validation routines provided on the CD-ROM
included with this book. These -- and other -- routines are more completely documented in
Chapter 13, "Plug-and-Play Routines." Also included here are various string processing routines
that are often helpful when working with forms.
Data Validation Routines
Input Processing Routines
Practical Examples of Input Validation
Here's an two example of how to use JavaScript for form input validation. Four text boxes are provided; beside each is a button. Enter text into one of the boxes, and click the button beside it. If you have not entered valid data, you are asked to re-enter it. Figure 16-3 shows the valid.htm document displayed in Netscape.
- Box 1 tests for a valid e-mail address. In the example the test is limited to merely looking for a @ symbol.
- Box 2 tests for input of exactly five characters.
- Box 3 tests for input of three or more characters.
- Box 4 tests for non-blank input.
valid.htm <CD> <HTML> <HEAD> <TITLE> Verifying Form Input with JavaScript</TITLE> <SCRIPT LANGUAGE="JavaScript"> function runTest(form, button) { Ret = false; if (button.name == "1") Ret = testBox1(form); if (button.name == "2") Ret = testBox2(form); if (button.name == "3") Ret = testBox3(form); if (button.name == "4") Ret = testBox4(form); if (Ret) alert ("Successful input!"); } function testBox1(form) { Ctrl = form.inputbox1; if (Ctrl.value == "" || Ctrl.value.indexOf ('@', 0) == -1) { validatePrompt (Ctrl, "Enter a valid email address") return (false); } else return (true); } function testBox2(form) { Ctrl = form.inputbox2; if (Ctrl.value.length != 5) { validatePrompt (Ctrl, "Provide five characters") return (false); } else return (true); } function testBox3(form) { Ctrl = form.inputbox3; if (Ctrl.value.length < 3) { validatePrompt (Ctrl, "Provide at least three characters") return (false); } else return (true); } function testBox4(form) { Ctrl = form.inputbox4; if (Ctrl.value == "") { validatePrompt (Ctrl, "Provide a value for this box") return (false); } else return (true); } function runSubmit (form, button) { if (!testBox1(form)) return; if (!testBox2(form)) return; if (!testBox3(form)) return; if (!testBox4(form)) return; alert ("All entries verified OK!"); //document.test.submit(); // un-comment to actually submit form return; } function validatePrompt (Ctrl, PromptStr) { alert (PromptStr) Ctrl.focus(); return; } </SCRIPT> </HEAD> <BODY> <FORM NAME="test" ACTION="" METHOD=GET> Enter an e-mail address (e.g. tj@myserver.com): <BR> <INPUT TYPE="text" NAME="inputbox1"> <INPUT TYPE="button" NAME="1" VALUE="Test Input" onClick="runTest(this.form, this)"><P> Enter five characters only: <BR> <INPUT TYPE="text" NAME="inputbox2"> <INPUT TYPE="button" NAME="2" VALUE="Test Input" onClick="runTest(this.form, this)"><P> Enter three or more characters: <BR> <INPUT TYPE="text" NAME="inputbox3"> <INPUT TYPE="button" NAME="3" VALUE="Test Input" onClick="runTest(this.form, this)"><P> Enter anything (don't leave blank): <BR> <INPUT TYPE="text" NAME="inputbox4"> <INPUT TYPE="button" NAME="4" VALUE="Test Input" onClick="runTest(this.form, this)"><P><P> <INPUT TYPE="button" NAME="Submit" VALUE="Submit" onClick="runSubmit(this.form, this)"><P> </FORM> </BODY> </HTML>
<Figure 16-3. The valid.htm document, ready for text input.>
Experiment with the valid.htm to see how the various validation routines work. The example actually double-checks each input; then checks again when you click the Submit button. In actual use, you need only test once, typically when the Submit button is pressed. When all the blanks have been entered correctly the script announces " All entries verified OK!" The focus method is used in the above example to set focus in a text box to prompt the user that the input for that box is incorrect. Under some platforms (such as Windows) the focus() method doesn't always display a flashing bar when the insertion point is placed in a text box. Focus is properly set in the desired box, but there may not be visual confirmation of it. Therefore, don't assume the user is aware that focus has been returned to the box. If your form contains many input boxes, specify the box that needs correct input by name.
Validating Non-text Controls
The text control is the natural candidate for form validation (hidden text controls need no user validation, and in Netscape 2.0 the value of the password text control cannot be read). In the previous examples you've seen how to validate the content of text controls. Here's how to validate other commonly use controls and control structures in HTML forms.
In all of the following, the form name is myform, and the control name is control. A typical example:
<SCRIPT LANGUAGE="JavaScript"> function testform() { alert (document.myform.control.value) } </SCRIPT> <FORM NAME="myform"> <TEXTAREA NAME="control" COLS=40 ROWS=5> </TEXTAREA> <INPUT TYPE="button" VALUE="Click" onClick="testform()"> </FORM>
Validating Textareas
As with text boxes, use the value property to test the content of textareas. Examples:
Ret = document.myform.control.value; // assigns content to Ret Ret = document.myform.control.value.length; // assigns length of content to Ret Ret = document.myform.control.value.indexOf ("\n") // value 0 or above indicates hard return
Validating Check boxes
The check box controls are either on or off. Use the checked property of the control to validate if a checkbox is checked or not checked. Examples:
Ret = document.myform.control.checked; // true if checked; false if not checked
Validating Radio Buttons
Only one radio button in a group can be selected at one time. You can check which one is selected using a loop like the following (this is described earlier in the chapter; it checks three buttons in the rad group):
for (Count = 0; Count < 3; Count++) { if (form.rad[Count].checked) break; }
You can use this loop to determine if none of the radio buttons are selected. This may be needed if you must initially display all of the radio buttons in a group as unselected, but require one of them to be selected.
Selected = false; for (Count = 0; Count < 3; Count++) { if (form.rad[Count].checked) { Selected = true; break; } } if (Selected) // a radio button is selected else // no radio button is selected
Validating Selection Lists
You may validate that at least one option in a selection list is selected with the following for loop:
Selected = false; for (Count = 0; Count < document.myform.control.length; Count++) { if (form.list[Count].selected) Selected = true; } if (Selected) // an option is selected else // no option is selected
When using multiple-choice selection lists you can verify that at least a certain number of options are selected with the following (this assumes at least three options in the list must be selected to pass verification):
Selected = 0; for (Count = 0; Count < document.myform.control.length; Count++) { if (form.list[Count].selected) Selected++; } if (Selected < 3) // fewer than 3 selected else // 3 or more selected
You can use opposite logic to verify that no more than a certain number of options are selected:
Selected = 0; for (Count = 0; Count < document.myform.control.length; Count++) { if (form.list[Count].selected) Selected++; } if (Selected > 3) // more than three selected else // 3 or less selected
Revised: October 31, 1996
URL: | http://www.webreference.com/content/jssource/chap16.html | CC-MAIN-2017-04 | refinedweb | 3,931 | 63.39 |
Learn how to set up Next.js with commonly used testing tools: Cypress, Playwright, and Jest with React Testing Library.
Cypress is a test runner used for End-to-End (E2E) and Integration Testing.
You can use
create-next-app with the with-cypress example to quickly get started.
npx create-next-app@latest --example with-cypress with-cypress-app
To get started with Cypress, install the
cypress package:
npm install --save-dev cypress
Add Cypress to the
package.json scripts field:
"scripts": { "dev": "next dev", "build": "next build", "start": "next start", "cypress": "cypress open", }
Run Cypress for the first time to generate examples that use their recommended folder structure:
npm run cypress
You can look through the generated examples and the Writing Your First Test section of the Cypress Documentation to help you get familiar with Cypress.
Assuming the following two Next.js pages:
// pages/index.js import Link from 'next/link' export default function Home() { return ( <nav> <Link href="/about"> <a>About</a> </Link> </nav> ) }
// pages/about.js export default function About() { return ( <div> <h1>About Page</h1> </div> ) }
Add a test to check your navigation is working correctly:
// cypress/integration/app.spec.js describe('Navigation', () => { it('should navigate to the about page', () => { // Start from the index page cy.visit('') // Find a link with an href attribute containing "about" and click it cy.get('a[href*="about"]').click() // The new url should include "/about" cy.url().should('include', '/about') // The new page should contain an h1 with "About page" cy.get('h1').contains('About Page') }) })
You can use
cy.visit("/") instead of
cy.visit("") if you add
"baseUrl": "" to the
cypress.json configuration file.
Since Cypress is testing a real Next.js application, it requires the Next.js server to be running prior to starting Cypress. We recommend running your tests against your production code to more closely resemble how your application will behave.
Run
npm run build and
npm run start, then run
npm run cypress in another terminal window to start Cypress.
Note: Alternatively, you can install the
start-server-and-testpackage and add it to the
package.jsonscripts field:
"test": "start-server-and-test start cypress"to start the Next.js production server in conjunction with Cypress. Remember to rebuild your application after new changes.
You will have noticed that running Cypress so far has opened an interactive browser which is not ideal for CI environments. You can also run Cypress headlessly using the
cypress run command:
// package.json "scripts": { //... "cypress": "cypress open", "cypress:headless": "cypress run", "e2e": "start-server-and-test start cypress", "e2e:headless": "start-server-and-test start cypress:headless" }
You can learn more about Cypress and Continuous Integration from these resources:
Playwright is a testing framework that lets you automate Chromium, Firefox, and WebKit with a single API. You can use it to write End-to-End (E2E) and Integration tests across all platforms.
The fastest way to get started is to use
create-next-app with the with-playwright example. This will create a Next.js project complete with Playwright all set up.
npx create-next-app@latest --example with-playwright with-playwright-app
You can also use
npm init playwright to add Playwright to an existing
NPM project.
To manually get started with Playwright, install the
@playwright/test package:
npm install --save-dev @playwright/test
Add Playwright to the
package.json scripts field:
"scripts": { "dev": "next dev", "build": "next build", "start": "next start", "test:e2e": "playwright test", }
Assuming the following two Next.js pages:
// pages/index.js import Link from 'next/link' export default function Home() { return ( <nav> <Link href="/about"> <a>About</a> </Link> </nav> ) }
// pages/about.js export default function About() { return ( <div> <h1>About Page</h1> </div> ) }
Add a test to verify that your navigation is working correctly:
// e2e/example.spec.ts import { test, expect } from '@playwright/test' test('should navigate to the about page', async ({ page }) => { // Start from the index page (the baseURL is set via the webServer in the playwright.config.ts) await page.goto('') // Find an element with the text 'About Page' and click on it await page.click('text=About') // The new URL should be "/about" (baseURL is used there) await expect(page).toHaveURL('') // The new page should contain an h1 with "About Page" await expect(page.locator('h1')).toContainText('About Page') })
You can use
page.goto("/") instead of
page.goto(""), if you add
"baseURL": "" to the
playwright.config.ts configuration file.
Since Playwright is testing a real Next.js application, it requires the Next.js server to be running prior to starting Playwright. It is recommended to run your tests against your production code to more closely resemble how your application will behave.
Run
npm run build and
npm run start, then run
npm run test:e2e in another terminal window to run the Playwright tests.
Note: Alternatively, you can use the
webServerfeature to let Playwright start the development server and wait until it's fully available.
Playwright will by default run your tests in the headless mode. To install all the Playwright dependencies, run
npx playwright install-deps.
You can learn more about Playwright and Continuous Integration from these resources:
Jest and React Testing Library are frequently used together for Unit Testing. There are three ways you can start using Jest within your Next.js application:
The following sections will go through how you can set up Jest with each of these options:
You can use
create-next-app with the with-jest example to quickly get started with Jest and React Testing Library:
npx create-next-app@latest --example with-jest with-jest-app
Since the release of Next.js 12, Next.js now has built-in configuration for Jest.
To set up Jest, install
jest,
jest-environment-jsdom,
@testing-library/react,
@testing-library/jest-dom:
npm install --save-dev jest jest-environment-jsdom @testing-library/react @testing-library/jest-dom
Create a
jest.config.js file in your project's root directory and add the following:
// jest.config.js const nextJest = require('next/jest') const createJestConfig = nextJest({ // Provide the path to your Next.js app to load next.config.js and .env files in your test environment dir: './', }) // Add any custom config to be passed to Jest /** @type {import('jest').Config} */ const customJestConfig = { // Add more setup options before each test is run // setupFilesAfterEnv: ['<rootDir>/jest.setup.js'], // if using TypeScript with a baseUrl set to the root directory then you need the below for alias' to work moduleDirectories: ['node_modules', '<rootDir>/'], testEnvironment: 'jest-environment-jsdom', } // createJestConfig is exported this way to ensure that next/jest can load the Next.js config which is async module.exports = createJestConfig(customJestConfig)
Under the hood,
next/jest is automatically configuring Jest for you, including:
transformusing SWC
.css,
.module.css, and their scss variants) and image imports
.env(and all variants) into
process.env
node_modulesfrom test resolving and transforms
.nextfrom test resolving
next.config.jsfor flags that enable SWC transforms
Note: To test environment variables directly, load them manually in a separate setup script or in your
jest.config.jsfile. For more information, please see Test Environment Variables.
If you opt out of the Rust Compiler, you will need to manually configure Jest and install
babel-jest and
identity-obj-proxy in addition to the packages above.
Here are the recommended options to configure Jest for Next.js:
// jest.config.js module.exports = { collectCoverage: true, // on node 14.x coverage provider v8 offers good speed and more or less good report coverageProvider: 'v8', collectCoverageFrom: [ '**/*.{js,jsx,ts,tsx}', '!**/*.d.ts', '!**/node_modules/**', '!<rootDir>/out/**', '!<rootDir>/.next/**', '!<rootDir>/*.config.js', '!<rootDir>/coverage/**', ], moduleNameMapper: { // Handle CSS imports (with CSS modules) // '^.+\\.module\\.(css|sass|scss)$': 'identity-obj-proxy', // Handle CSS imports (without CSS modules) '^.+\\.(css|sass|scss)$': '<rootDir>/__mocks__/styleMock.js', // Handle image imports // '^.+\\.(png|jpg|jpeg|gif|webp|avif|ico|bmp|svg)$/i': `<rootDir>/__mocks__/fileMock.js`, // Handle module aliases '^@/components/(.*)$': '<rootDir>/components/$1', }, // Add more setup options before each test is run // setupFilesAfterEnv: ['<rootDir>/jest.setup.js'], testPathIgnorePatterns: ['<rootDir>/node_modules/', '<rootDir>/.next/'], testEnvironment: 'jsdom', transform: { // Use babel-jest to transpile tests with the next/babel preset // '^.+\\.(js|jsx|ts|tsx)$': ['babel-jest', { presets: ['next/babel'] }], }, transformIgnorePatterns: [ '/node_modules/', '^.+\\.module\\.(css|sass|scss)$', ], }
You can learn more about each configuration option in the Jest docs.
Handling stylesheets and image imports
Stylesheets and images aren't used in the tests but importing them may cause errors, so they will need to be mocked. Create the mock files referenced in the configuration above -
fileMock.js and
styleMock.js - inside a
__mocks__ directory:
// __mocks__/fileMock.js module.exports = { src: '/img.jpg', height: 24, width: 24, blurDataURL: ', }
// __mocks__/styleMock.js module.exports = {}
For more information on handling static assets, please refer to the Jest Docs.
Optional: Extend Jest with custom matchers
@testing-library/jest-dom includes a set of convenient custom matchers such as
.toBeInTheDocument() making it easier to write tests. You can import the custom matchers for every test by adding the following option to the Jest configuration file:
// jest.config.js setupFilesAfterEnv: ['<rootDir>/jest.setup.js']
Then, inside
jest.setup.js, add the following import:
// jest.setup.js import '@testing-library/jest-dom/extend-expect'
If you need to add more setup options before each test, it's common to add them to the
jest.setup.js file above.
Optional: Absolute Imports and Module Path Aliases
If your project is using Module Path Aliases, you will need to configure Jest to resolve the imports by matching the paths option in the
jsconfig.json file with the
moduleNameMapper option in the
jest.config.js file. For example:
// tsconfig.json or jsconfig.json { "compilerOptions": { "baseUrl": ".", "paths": { "@/components/*": ["components/*"] } } }
// jest.config.js moduleNameMapper: { '^@/components/(.*)$': '<rootDir>/components/$1', }
Add a test script to package.json
Add the Jest executable in watch mode to the
package.json scripts:
"scripts": { "dev": "next dev", "build": "next build", "start": "next start", "test": "jest --watch" }
jest --watch will re-run tests when a file is changed. For more Jest CLI options, please refer to the Jest Docs.
Create your first tests
Your project is now ready to run tests. Follow Jest's convention by adding tests to the
__tests__ folder in your project's root directory.
For example, we can add a test to check if the
<Home /> component successfully renders a heading:
// __tests__/index.test.jsx import { render, screen } from '@testing-library/react' import Home from '../pages/index' import '@testing-library/jest-dom' describe('Home', () => { it('renders a heading', () => { render(<Home />) const heading = screen.getByRole('heading', { name: /welcome to next\.js!/i, }) expect(heading).toBeInTheDocument() }) })
Optionally, add a snapshot test to keep track of any unexpected changes to your
<Home /> component:
// __tests__/snapshot.js import { render } from '@testing-library/react' import Home from '../pages/index' it('renders homepage unchanged', () => { const { container } = render(<Home />) expect(container).toMatchSnapshot() })
Note: Test files should not be included inside the pages directory because any files inside the pages directory are considered routes.
Running your test suite
Run
npm run test to run your test suite. After your tests pass or fail, you will notice a list of interactive Jest commands that will be helpful as you add more tests.
For further reading, you may find these resources helpful:
The Next.js community has created packages and articles you may find helpful:
For more information on what to read next, we recommend: | https://nextjs.org/docs/testing | CC-MAIN-2022-40 | refinedweb | 1,888 | 50.73 |
The attached patches introduce a couple of new dentry operations for use by automounters and make AFS, NFS, CIFS and autofs4 use them. This means that these filesystems no longer have to abuse lookup(), follow_link() and d_revalidate() to achieve the desired effect. There are two dentry operations provided: (1) struct vfsmount *(*d_automount)(struct path *path); This is used by follow_automount() in fs/namei.c to ask the filesystem that owns the dentry at the current path point to mount something on @path. It is called if either the inode belonging to the given dentry is flagged S_AUTOMOUNT or the dentry is flagged DMANAGED_AUTOMOUNT, and if the dentry has nothing mounted on it in the current namespace when someone attempts to use that dentry. No locks will be held when this is called. d_op->d_automount() may return one of: (a) The vfsmount mounted upon that dentry, in which case pathwalk will move to the root dentry of that vfsmount. (b) NULL if something was already mounted there, in which case pathwalk will loop around and recheck the mountings. (c) -EISDIR, in which case pathwalk will stop at this point and attempt to use that dentry as the object of interest. If the current dentry is not terminal within the path, -EREMOTE will be returned. (d) An error value, to be returned immediately. Automount transits are counted as symlinks to prevent circular references from being a problem. If one is detected, -ELOOP will be returned. If stat() is given AT_NO_AUTOMOUNT then d_op->d_automount() will not be invoked on a terminal dentry; instead that dentry will be returned by pathwalk. follow_automount() also does not invoke d_op->d_automount() if the caller gave AT_SYMLINK_NOFOLLOW to stat(), but rather returns the base dentry. (2) int (*d_manage)(struct path *path, bool mounting_here); This is called by managed_dentry() or follow_down() in fs/namei.c to indicate to a filesystem that pathwalk is about to step off of the current path point and walk to another point in the path. This is called if DMANAGED_TRANSIT is set on a dentry. This can then be used by autofs to stop non-daemon processes from walking until it has finished constructing or expiring the tree behind the dentry. It could also be used to prevent undesirables from mounting on this dentry. @mounting_here is true if called from follow_down() from mount, in which case namespace_sem is held exclusively by the caller of follow_down(). Otherwise, no locks are held. d_op->d_manage() may return one of: (a) 0 to continue the pathwalk as normal. (b) -EISDIR to prevent managed_dentry() from crossing to a mounted filesystem or calling d_op->d_automount(), in which case the dentry will be treated as an ordinary directory. (c) An error to abort the pathwalk completely. To make this work for autofs, d_mounted in struct dentry has become d_managed. This used to be a count of the number of mounts on this dentry. The bottom 28 bits are still that (DMANAGED_MOUNTPOINT). The upper four bits contain a couple of flags, DMANAGED_TRANSIT and DMANAGED_AUTOMOUNT. This allows managed_dentry() to test all three conditions with minimumal overhead if none of them are true. For other filesystems, setting S_AUTOMOUNT is sufficient. This is noted by __d_instantiate() which will set DMANAGED_AUTOMOUNT automatically if it is seen. Checking S_AUTOMOUNT doesn't work for autofs, however, since the dentry might not have an inode, hence why a dentry flag also. S_AUTOMOUNT and d_automount() are introduced in patch 1; d_manage(), d_managed and DMANAGED_* are introduced in patch 7. David --- David Howells (8): Make follow_down() handle d_manage() Make dentry::d_mounted into a more general field for special function dirs Add an AT_NO_AUTOMOUNT flag to suppress terminal automount Remove the automount through follow_link() kludge code from pathwalk CIFS: Use d_automount() rather than abusing follow_link() NFS: Use d_automount() rather than abusing follow_link() AFS: Use d_automount() rather than abusing follow_link() Add a dentry op to handle automounting rather than abusing follow_link() Ian Kent (9): autofs4 - bump version autofs4 - add v4 pseudo direct mount support autofs4 - fix wait validation autofs4: cleanup autofs4_free_ino() autofs4: cleanup dentry operations autofs4: cleanup inode operations autofs4: removed unused code autofs4: add d_manage() dentry operation autofs4: add d_automount() dentry operation Documentation/filesystems/Locking | 2 Documentation/filesystems/vfs.txt | 22 + fs/afs/dir.c | 1 fs/afs/inode.c | 3 fs/afs/internal.h | 1 fs/afs/mntpt.c | 47 +-- fs/autofs/dirhash.c | 5 fs/autofs4/autofs_i.h | 100 ++++-- fs/autofs4/dev-ioctl.c | 2 fs/autofs4/expire.c | 42 ++ fs/autofs4/inode.c | 28 -- fs/autofs4/root.c | 668 ++++++++++++++++--------------------- fs/autofs4/waitq.c | 17 + fs/cifs/cifs_dfs_ref.c | 134 ++++--- fs/cifs/cifsfs.h | 6 fs/cifs/dir.c | 2 fs/cifs/inode.c | 8 fs/dcache.c | 7 fs/namei.c | 243 +++++++++++-- fs/namespace.c | 20 + fs/nfs/dir.c | 4 fs/nfs/inode.c | 4 fs/nfs/internal.h | 1 fs/nfs/namespace.c | 87 ++--- fs/nfsd/vfs.c | 5 fs/stat.c | 4 include/linux/auto_fs4.h | 2 include/linux/dcache.h | 19 + include/linux/fcntl.h | 1 include/linux/fs.h | 2 include/linux/namei.h | 5 include/linux/nfs_fs.h | 1 32 files changed, 836 insertions(+), 657 | https://lwn.net/Articles/407940/ | CC-MAIN-2017-30 | refinedweb | 852 | 51.55 |
I've spent two whole days trying to create an initial migration for the database of my project. This is so fustrating. Each preview version of the docs points towards different directions, and there're a lot of unclosed issues flying arround for a while.
My project is an AspNetCore application running on the full framework (net462) although I think I've tryed every combination of preview versions, even the workarounds proposed on this issue: EF Tools 1.1.0-preview4 Unrecognized option '--config' or in this one: but neither work.
This is an abstract of my project.json with the relevant parts:
{ "version": "1.0.0-*", "buildOptions": { "platform": "x86", "debugType": "full", "preserveCompilationContext": true, "emitEntryPoint": true }, "dependencies": { .... "Microsoft.EntityFrameworkCore": "1.1.0", "Microsoft.EntityFrameworkCore.Design": "1.1.0", "Microsoft.EntityFrameworkCore.SqlServer": "1.1.0", "Microsoft.EntityFrameworkCore.SqlServer.Design": "1.1.0", .... }, "tools": { "Microsoft.AspNetCore.Server.IISIntegration.Tools": "1.1.0-preview4-final", "Microsoft.EntityFrameworkCore.Tools.DotNet": "1.1.0-preview4-final" }, "frameworks": { "net462": { } }, ... }
In my case the proposed workarounds don't work, neither using the nightly builds nor downgrading the tools to 1.0.0-preview3.
If I use the 1.1.0-preview4-final version of the tools I hit this error:
Unrecognized option --config
If I use the nightly builds I get this one, wich is somehow absurd, as my app has only one project and is not a dll (it has also emitEntryPoint:true set)
Could not load assembly 'Sales'. Ensure it is referenced by the startup project 'Sales'
But this is my favourite one, when I downgrade to the 1.0.0-preview3-final of the tools I get this surrealistic one:
error: Package Microsoft.EntityFrameworkCore.Tools.DotNet 1.0.0-preview3-final is not compatible with netcoreapp1.0 (.NETCoreApp,Version=v1.0). Package Microsoft.EntityFrameworkCore.Tools.DotNet 1.0.0-preview3-final supports: netcoreapp1.0 (.NETCoreApp,Version=v1.0)
I had to read it five times to get sure that in the second sentence was telling just the opposite of the first one... It seems a joke!
Furthermore, commands are not working on the PMC anymore, no matter wich version of the tools I install, no matter if I restore the packages and if I restart the computer...
I'm getting crazy with so many versions of everything and I only want to create a migration, it doesn't matter wich version of the tools I have to use... Is there a valid configuration nowadays or am I trying something imposible? Has anybody been able to create migrations within an asp.net core application targeting the full .net framework (net462) with ANY version of the ef tooling?
If so, HOW?
EDIT:
After targeting the project to .netcoreapp1.0 and removing the incompatible references now I hit this error:
A fatal error was encountered. The library 'hostpolicy.dll' required to execute the application was not found in 'C:\Program Files (x86)\dotnet\shared\Microsoft.NETCore.App\1.0.1'
What's happening here??? I'm really tired of .net Core, and it's still in it's first version. I've suffered a lot of issues like this while it was in beta, but now things are supposed to be stable... They have changed twenty times everything that can be changed, APIs, assembly names, namespaces, package names, conventions... Now let's wait for the preview5, 6 or 25 of the tooling and maybe by the year 2035 EF Core will have appropiate tools and procedures, meanwhile I damn a million time my decission of betting for this technology!
EDIT 2:
As per comments global.json may be relevant:
{ "projects": [ "src", "test" ], "sdk": { "version": "1.0.0-preview2-1-003177" } }
and to add that the 1.0.0-preview2-1-003177 folder exists and is the only one in C:\Program Files (x86)\dotnet\sdk\ and C:\Program Files\dotnet\sdk\
I hate to answer my own question, but I suppose that not too much people will go into this alley... So for those who are struggling with a similar problem I'll tell that mine came from this configuration on project.json:
... "buildOptions": { "platform": "x86", <- THIS!!! "debugType": "portable", "preserveCompilationContext": true, "emitEntryPoint": true },
after removing the "platform" key, migrations started to work again...
I'm not really sure when did I introduced that setting, since I didn't try to create migrations before upgrading to the version 1.1 of the .NET Core SDK. Maybe it was copied from one of the examples on internet, maybe it was from a previous version, I don't know, but that has turned me crazy for days, I hope it helps somebody outthere. | https://entityframeworkcore.com/knowledge-base/41068199/can-t-create-a-migration-with-ef-core-1-1 | CC-MAIN-2020-34 | refinedweb | 773 | 57.06 |
Red Hat Bugzilla – Bug 1251423
Should add asterisk to all required fields on creation page.
Last modified: 2015-11-23 16:15:55 EST
Description of problem:
When create app from source code or from template, on creation page, there is no asterisk to highlight the parameter that must be defined. Such as the app name is required but not highlighted on creation page from source code.
Version-Release number of selected component (if applicable):
devenv_fedora_2115
$ oc version
oc v1.0.4-88-ga2ad7d7
kubernetes v1.0.0
How reproducible:
Always
Steps to Reproduce:
1.Create images in openshift project
$ oc create -f /data/src/github.com/openshift/origin/examples/image-streams/image-streams-centos7.json -n openshift
2.Create templates in project.
$ oc create -f /data/src/github.com/openshift/origin/examples/sample-app/application-template-stibuild.json -n protest
3.On web console, when create app using the template， check the creation page.
4.On web console, when create app using source code, check the creation page.
Actual results:
3.On the creation page, there are many parameters required, but no asterisk to highlight them. User would forget to define some parameters.
4 On the creation page, there is no asterisk to prompt the parameter is required.
Expected results:
3,4.Should have asterisk to prompt user the parameter is required.
Additional info:
We've previously added an asterisk for required template parameters:
Can you try in a newer build?
Required template parameters are done.
Name and Replicas are not yet marked as required on the last create from source page.
Checked on devenv_fedora_2129, the code from pull 3986 has been merged on the test env, but still no asterisk found on creation page. Could you help check what's the issue, Samuel?
Created attachment 1061015 [details]
Modified ruby hello world template
The parameter needs to be marked as required in the template. Can you check?
A required parameter looks like
{
"name": "MYSQL_DATABASE",
"description": "database name",
"value": "",
"required": true
}
You can see with the command
oc get template <name> -o json -n <namespace>
I've also attached a template with required parameters.
Test with the template with required parameters, and there are asterisks on creation page now. So example templates under origin/example/ need to be updated accordingly.
(In reply to Yanping Zhang from comment #6)
> Test with the template with required parameters, and there are asterisks on
> creation page now. So example templates under origin/example/ need to be
> updated accordingly.
Ben ^^
Sam are you just asking that i go update fields in the existing templates to add the "required=true" flag?
Ben, not necessarily. Just making you aware of this bug and seeing if we plan to add required to those templates.
I didn't have any plans to do so.. not that it would be necessarily wrong to, but the real use case for "required" is a field that is by default empty, but needs a value to be provided.
in our db templates, the fields need a value, but we provide a default one. So unless the user goes in and explicitly deletes/overrides the default value with an empty string, they won't have problems.
I'm concerned about the user experience when you do happen to remove that value, though. Only some of the template resources are created, and you have to figure out how to delete things and start over. It's very easy to catch before submit if required is set.
The resources all get created. but the DB pod (for example) will fail to start citing the lack of a DB_NAME/USER/PASSWORD.
In a way that's worse because it looks like everything succeeded. Would a user know what's wrong and how to fix it?
It just seems like it could be easily avoided.
I believe the all required fields now have the asterisk.
Checked on devenv-fedora_2517.
openshift v1.0.6-833-ga7032bc
kubernetes v1.2.0-alpha.1-1107-g4c8e6f4
etcd 2.1.2
When create app on web console, on creation page, all required fields now have the asterisk.
Move the bug to Verified. | https://bugzilla.redhat.com/show_bug.cgi?id=1251423 | CC-MAIN-2017-43 | refinedweb | 690 | 59.4 |
WebKit Bugzilla
this would require to retrieve the JSGlobalContextRef and some functions as well as access to the required header files at compile time.
It sounds like this is basically the functionality that -[WebFrame windowObject], -[WebFrame globalContex] and the webView:didClearWindowObject:forFrame: delegate method available in the Mac port. This would seem to map to two WebKitFrame methods and a signal in the Gtk API.
I mailed Michael a patch that implements most of this functionality recently. It was blocked by a few minor issues. Are you able to complete it and attach it for review?
This looks like a tracking bug for a few distinct bugs for which I've added dependencies.
Created attachment 17639 [details]
Fix
Landed in r28313.
Until is fixed, applications will have to do:
#include <JavaScriptCore/JSBase.h>
#include <JavaScriptCore/JSContextRef.h>
#include <JavaScriptCore/JSStringRef.h>
#include <JavaScriptCore/JSObjectRef.h>
#include <JavaScriptCore/JSValueRef.h>
Comment on attachment 17639 [details]
Fix
Already r'ed by aroben, bug closed. | https://bugs.webkit.org/show_bug.cgi?id=15687 | CC-MAIN-2016-07 | refinedweb | 162 | 52.26 |
JavaScript Basics Before You Learn React
Nathan Sebhastian
Jan 14
Updated on Feb 09, 2019
・9 min read
In an ideal world, you can learn all about JavaScript and web development before you dive into React. Unfortunately, we live in a not-perfect world, so chomping down on ALL of JavaScript before React will just make you bleed hard. If you already have some experience with JavaScript, all you need to learn before React is just the JavaScript features you will actually use to develop React application. Things about JavaScript you should be comfortable with before learning React are:
- ES6 classes
- The new variable declaration let/const
- Arrow functions
- Destructuring assignment
- Map and filter
- ES6 module system
It's the 20% of JavaScript features that you will use 80% of the time, so in this tutorial I will help you learn them all.
Exploring Create React App
The usual case of starting to learn React is to run the
create-react-app package, which sets up everything you need to run React. Then after the process is finished, opening
src/app.js will present us with the only React class in the whole app:;
If you never learned ES6 before, you'd think that this class statement is a feature of React. It's actually a new feature of ES6, and that's why learning ES6 properly would enable you to understand React code better. We'll start with ES6 classes.
ES6 Classes
ES6 introduced class syntax that is used in similar ways to OO language like Java or Python. A basic class in ES6 would look like this:
class Developer { constructor(name){ this.name = name; } hello(){ return 'Hello World! I am ' + this.name + ' and I am a web developer'; } }
class syntax is followed by an identifier (or simply name) that can be used to create new objects. The
constructor method is always called in object initialization. Any parameters passed into the object will be passed into the new object. For example:
var nathan = new Developer('Nathan'); nathan.hello(); // Hello World! I am Nathan and I am a web developer
A class can define as many methods as the requirements needed, and in this case, we have the
hello method which returns a string.
Class inheritance
A class can
extends the definition of another class, and a new object initialized from that class will have all the methods of both classes.
class ReactDeveloper extends Developer { installReact(){ return 'installing React .. Done.'; } } var nathan = new ReactDeveloper('Nathan'); nathan.hello(); // Hello World! I am Nathan and I am a web developer nathan.installReact(); // installing React .. Done.
The class that
extends another class is usually called child class or sub class, and the class that is being extended is called parent class or super class. A child class can also override the methods defined in parent class, meaning it will replace the method definition with the new method defined. For example, let's override the
hello function:
class ReactDeveloper extends Developer { installReact(){ return 'installing React .. Done.'; } hello(){ return 'Hello World! I am ' + this.name + ' and I am a REACT developer'; } } var nathan = new ReactDeveloper('Nathan'); nathan.hello(); // Hello World! I am Nathan and I am a REACT developer
There you go. The
hello method from
Developer class has been overridden.
Use in React
Now that we understand ES6 class and inheritance, we can understand the React class defined in
src/app.js. This is a React component, but it's actually just a normal ES6 class which inherits the definition of React Component class, which is imported from the React package.
import React, { Component } from 'react'; class App extends Component { // class content render(){ return ( <h1>Hello React!</h1> ) } }
This is what enables us to use the
render() method, JSX,
this.state, other methods. All of this definitions are inside the
Component class. But as we will see later, class is not the only way to define React Component. If you don't need state and other lifecycle methods, you can use a function instead.
Declaring variables with ES6
let and
const
Because JavaScript
var keyword declares variable globally, two new variable declarations were introduced in ES6 to solve the issue, namely
let and
const. They are all the same, in which they are used to declare variables. The difference is that
const cannot change its value after declaration, while
let can. Both declarations are local, meaning if you declare
let inside a function scope, you can't call it outside of the function.
const name = "David"; let age = 28; var occupation = "Software Engineer";
Which one to use?
The rule of thumb is that declare variable using
const by default. Later when you wrote the application, you'll realize that the value of
const need to change. That's the time you should refactor
const into
let. Hopefully it will make you get used to the new keywords, and you'll start to recognize the pattern in your application where you need to use
const or
let.
When do we use it in React?
Everytime we need variables. Consider the following example:
import React, { Component } from 'react'; class App extends Component { // class content render(){ const greeting = 'Welcome to React'; return ( <h1>{greeting}</h1> ) } }
Since greeting won't change in the entire application lifecycle, we define it using
const here.
The arrow function
Arrow function is a new ES6 feature that's been used almost widely in modern codebases because it keeps the code concise and readable. This feature allows us to write functions using shorter syntax
// regular function const testFunction = function() { // content.. } // arrow function const testFunction = () => { // content.. }
If you're an experienced JS developer, moving from the regular function syntax to arrow syntax might be uncomfortable at first. When I was learning about arrow function, I used this simple 2 steps to rewrite my functions:
- remove function keyword
- add the fat arrow symbol
=>after
()
the parentheses are still used for passing parameters, and if you only have one parameter, you can omit the parentheses.
const testFunction = (firstName, lastName) => { return firstName+' '+lastName; } const singleParam = firstName => { return firstName; }
Implicit return
If your arrow function is only one line, you can return values without having to use the
return keyword and the curly brackets
{}
const testFunction = () => 'hello there.'; testFunction();
Use in React
Another way to create React component is to use arrow function. React take arrow function:
const HelloWorld = (props) => { return <h1>{props.hello}</h1>; }
as equivalent to an ES6 class component
class HelloWorld extends Component { render() { return ( <h1>{props.hello}</h1>; ); } }
Using arrow function in your React application makes the code more concise. But it will also remove the use of state from your component. This type of component is known as stateless functional component. You'll find that name in many React tutorials.
Destructuring assignment for arrays and objects
One of the most useful new syntax introduced in ES6, destructuring assignment is simply copying a part of object or array and put them into named variables. A quick example:
const developer = { firstName: 'Nathan', lastName: 'Sebhastian', developer: true, age: 25, } //destructure developer object const { firstName, lastName } = developer; console.log(firstName); // returns 'Nathan' console.log(lastName); // returns 'Sebhastian' console.log(developer); // returns the object
As you can see, we assigned firstName and lastName from
developer object into new variable
firstName and
lastName. Now what if you want to put
firstName into a new variable called
name?
const { firstName:name } = developer; console.log(name); // returns 'Nathan'
Destructuring also works on arrays, only it uses index instead of object keys:
const numbers = [1,2,3,4,5]; const [one, two] = numbers; // one = 1, two = 2
You can skip some index from destructuring by passing it with
,:
const [one, two, , four] = numbers; // one = 1, two = 2, four = 4
Use in React
Mostly used in destructuring
state in methods, for example:
reactFunction = () => { const { name, email } = this.state; };
Or in functional stateless component, consider the example from previous chapter:
const HelloWorld = (props) => { return <h1>{props.hello}</h1>; }
We can simply destructure the parameter immediately:
const HelloWorld = ({ hello }) => { return <h1>{hello}</h1>; }
Map and filter
Although this tutorial focuses on ES6, JavaScript array
map and
filter methods need to be mentioned since they are probably one of the most used ES5 features when building React application. Particularly on processing data.
These two methods are much more used in processing data. For example, imagine a fetch from API result returns an array of JSON data:
const users = [ { name: 'Nathan', age: 25 }, { name: 'Jack', age: 30 }, { name: 'Joe', age: 28 }, ];
Then we can render a list of items in React as follows:
import React, { Component } from 'react'; class App extends Component { // class content render(){ const users = [ { name: 'Nathan', age: 25 }, { name: 'Jack', age: 30 }, { name: 'Joe', age: 28 }, ]; return ( <ul> {users .map(user => <li>{user.name}</li>) } </ul> ) } }
We can also filter the data in the render.
<ul> {users .filter(user => user.age > 26) .map(user => <li>{user.name}</li>) } </ul>
ES6 module system
The ES6 module system enables JavaScript to import and export files. Let's see the
src/app.js code again in order to explain this.;
Up at the first line of code we see the import statement:
import React, { Component } from 'react';
and at the last line we see the
export default statement:
export default App;
To understand these statements, let's discuss about modules syntax first.
A module is simply a JavaScript file that exports one or more values (can be objects, functions or variables) using the
export keyword. First, create a new file named
util.js in the
src directory
touch util.js
Then write a function inside it. This is a default export
export default function times(x) { return x * x; }
or multiple named exports
export function times(x) { return x * x; } export function plusTwo(number) { return number + 2; }
Then we can import it from
src/App.js
import { times, plusTwo } from './util.js'; console.log(times(2)); console.log(plusTwo(3));
You can have multiple named exports per module but only one default export. A default export can be imported without using the curly braces and corresponding exported function name:
// in util.js export default function times(x) { return x * x; } // in app.js import k from './util.js'; console.log(k(4)); // returns 16
But for named exports, you must import using curly braces and the exact name. Alternatively, imports can use alias to avoid having the same name for two different imports:
// in util.js export function times(x) { return x * x; } export function plusTwo(number) { return number + 2; } // in app.js import { times as multiplication, plusTwo as plus2 } from './util.js';
Import from absolute name like:
import React from 'react';
Will make JavaScript check on
node_modules for the corresponding package name. So if you're importing a local file, don't forget to use the right path.
Use in React
Obviously we've seen this in the
src/App.js file, and then in
index.js file where the exported
App component is being rendered. Let's ignore the serviceWorker part for now.
//index.js file();
Notice how App is imported from
./App directory and the
.js extension has been omitted. We can leave out file extension only when importing JavaScript files, but we have to include it on other files, such as
.css. We also import another node module
react-dom, which enables us to render React component into HTML element.
As for PWA, it's a feature to make React application works offline, but since it's disabled by default, there's no need to learn it in the beginning. It's better to learn PWA after you're confident enough building React user interfaces.
Conclusion
The great thing about React is that it doesn't add any foreign abstraction layer on top of JavaScript as other web frameworks. That's why React becomes very popular with JS developers. It simply uses the best of JavaScript to make building user interfaces easier and maintainable. There really is more of JavaScript than React specifix syntax inside a React application, so once you understand JavaScript better — particularly ES6 — you can write React application with confident. But it doesn't mean you have to master everything about JavaScript to start writing React app. Go and write one now, and as opportunities come your way, you will be a better developer.
And by the way, you might wanna check out this cool articles from Toptal:
Who's looking for open source contributors? (Dec 31st edition)
Find something to work on or promote your project here. Please shamelessly pro...
Arrow functions, apart from their aesthetics, have this property called
lexical scoping. This explains lexical scoping better than I ever will :)
In short, arrow functions follow the scope of the caller's scope, rather than having its own. function() {} functions' scope can change to whatever by calling .bind() - basically how JS prototype works.
Maybe you deliberately omitted this because in React world arrow functions are mostly used for shorter declaration. But hey, I figured the post might give a wrong impression that arrow function is only for cleaner code :P
Short test code:
`
Ah yes, sorry if I skip the part about this. Thanks for adding it here, Jesse :)
'Both declarations are local, meaning if you declare let inside a function scope, you can't call it outside of the function.'
a var declared inside a function is also not global. this is not what is different about let. the difference is that let has block scope for example:
let x = 1;
if (x === 1) {
let x = 2;
console.log(x);
// expected output: 2
}
console.log(x);
// expected output: 1
Oh sorry, what I actually meant is block scope not function scope. If you call x outside of the if scope it will return 2 with var. I'll fix that as soon possible. Thanks David :)
Great post Nathan. Easy to understand. Only thing i couldn't get my head around is the Destructuring concept coz i haven't used or heard it before. Hopefully i'll get better understanding after some use.
Thanks.
Thanks dan, don't worry too much about destructuring, in simple application, we used it for shortening the syntax for getting state value only. For example, if you have this state initialized
Then when you want to get the value, do:
Instead of:
Now in my example tutorial, I have also included destructuring assignment into a new variable, like:
But to tell you the truth, I never used this kind of assignment, so just consider it an extra knowledge that might be useful sometime 😆
yap...its super important to Understand JavaScript Fundamental before using any frameworks. One should learn how JavaScript works, its life cycle, What Prototypal inheritance is, Classical vs Prototypal model, What Closure is etc...
Thank you for this post! I am actually working on a new web project, and this time I decided to use React instead of JS vanilla. It helped me a lot :D
You're welcome Rospars. Glad could help you out. Good luck with your project :)
Great post, thanks Nathan.
Great selection of features. Great article! I would add the spread operator as well. It's useful for making shallow copies of arrays and objects.
I've used
{...this.props}to pass a copy of props in react without naming them all, and then using destructuring to extract them in the component. See this stackoverflow question for a good description.
Ah certainly Steve, spread and rest operator would be a great addition. I'm just afraid the article would be too long when I wrote this. Thanks for your comment :)
Could I translate your post into Korean? I'm working as a front-end developer but mainly maintaining the old legacy. Next version would use React. So, I'm studying React.
Sure, go ahead
Thanks :)
Thank you Nathan for a great and informative post
Good article for the beginner who want to learn reactJS.
Thanks Nathan.
Hey, great article, I knew most of this, but the part of modules have clarified me all I didn't have end up understanding.
BTW, I think there is an error here, no?
// in util.js
export default function times(x) {
return x * x;
}
// in app.js
export k from './util.js';
console.log(k(4)); // returns 16
That second export should be an import, shouldn't be?
ah yes, thanks for the correction Alex :) glad could help you learn something
Great post!
I'd like to also mention .forEach() and .reduce() which are not quite as commonly used as .map() and .filter() but still worth noting. :) | https://practicaldev-herokuapp-com.global.ssl.fastly.net/nsebhastian/javascript-basics-before-you-learn-react-38en | CC-MAIN-2019-09 | refinedweb | 2,765 | 64.3 |
Your Account
Display List: Chapter 6 - ActionScript 3.0 Cookbook
Pages: 1, 2, 3, 4
You want to change the order in which objects are drawn on-screen, moving them either in front of or behind other display objects.
Use the setChildIndex( ) method of the DisplayObectContainer class to change the position of a particular item. Use the getChildIndex( ) and getChildAt( ) methods to query siblings of the item so the item can be positioned properly relative to them.
setChildIndex( )
DisplayObectContainer
getChildIndex( )
getChildAt( )
Recipes 6.1 and 6.2 introduced how the display list model deals with the visual stacking order (depth). Essentially, every DisplayObjectContainer instance has a list of children, and the order of the children in this list determines the order in which child display objects are drawn inside of the container. The children are each given a position index, ranging from 0 to numChildren - 1, much like an array. The child at position 0 is drawn on the bottom, underneath the child at position 1, etc. There are no empty position values in the list; if there are three children, the children will always have index values of 0, 1, and 2 (and not, say, 0, 1, and 6).
DisplayObjectContainer
numChildren
- 1
The setChildIndex( ) method is provided by DisplayObjectContainer to reorder the children inside the container. It takes two parameters: a reference to the child to be moved and the child's new position in the container. The index position specified must be a valid value. Negative values or values too large will generate a RangeError and the function won't execute properly.
RangeError
The following example creates three colored circles, with the blue one being drawn on top. The setChildIndex( ) method is used to move the blue circle underneath the two other circles, changing its position from 2 to 0 in the container. The positions of the other children are adjusted accordingly; red is moved to 1 and green is moved to 2:
package {
import flash.display.*;
public class SetChildIndexExample extends Sprite {
public function SetChildIndexExample( ) {
// Create three different colored circles and
// change their coordinates so they are staggered
// and aren't all located at (0,0).
var red:Shape = createCircle( 0xFF0000, 10 );
red.x = 10;
red.y = 20;
var green:Shape = createCircle( 0x00FF00, 10 );
green.x = 15;
green.y = 25;
var blue:Shape = createCircle( 0x0000FF, 10 );
blue.x = 20;
blue.y = 20;
// Add the circles, red has index 0, green 1, and blue 2
addChild( red );
addChild( green );
addChild( blue );
// Move the blue circle underneath the others by placing
// it at the very bottom of the list, at index 0
setChildIndex( blue, 0 );
}
// Helper function to create a circle shape with a given color
// and radius
public function createCircle( color:uint, radius:Number ):Shape {
var shape:Shape = new Shape( );
shape.graphics.beginFill( color );
shape.graphics.drawCircle( 0, 0, radius );
shape.graphics.endFill( );
return shape;
}
}
}
One of the requirements for setChildIndex( ) is that you know the index value you want to give to a specific child. When you're sending a child to the back, you use 0 as the index. When you want to bring a child to the very front, you specify numChildren - 1 as the index. But what if you want to move a child underneath another child?
For example, suppose you have two circles--one green and one blue--and you don't know their positions ahead of time. You want to move the blue circle behind the green one, but setChildIndex( ) requires an integer value for the new position. There are no setChildAbove or setChildBelow methods, so the solution is to use the getChildIndex( ) method to retrieve the index of a child, and then use that index to change the position of the other child. The getChildIndex( ) method takes a display object as a parameter and returns the index of the display object in the container. If the display object passed in is not a child of the container, an ArgumentError is thrown.
setChildAbove
setChildBelow
ArgumentError
The following example creates two circles--one green and one blue--and uses getChildIndex( ) on the green circle so the blue circle can be moved beneath it. By setting the blue circle to the index that the green circle has, the blue circle takes over the position and the green circle moves to the next higher position because blue had a higher position initially:
package {
import flash.display.*;
public class GetChildIndexExample extends Sprite {
public function GetChildIndexExample( ) {
// Create two different sized circles
var green:Shape = createCircle( 0x00FF00, 10 );
green.x = 25;
green.y = 25;
var blue:Shape = createCircle( 0x0000FF, 20 );
blue.x = 25;
blue.y = 25;
// Add the circles to this container
addChild( green );
addChild( blue );
// Move the blue circle underneath the green circle. First
// the index of the green circle is retrieved, and then the
// blue circle is set to that index.
setChildIndex( blue, getChildIndex( green ) );
}
// Helper function to create a circle shape with a given color
// and radius
public function createCircle( color:uint, radius:Number ):Shape {
var shape:Shape = new Shape( );
shape.graphics.beginFill( color );
shape.graphics.drawCircle( 0, 0, radius );
shape.graphics.endFill( );
return shape;
}
}
}
When a child is moved to an index lower than the one it currently has, all children from the target index up to the one just before the child index will have their indexes increased by 1 and the child is assigned to the target index. When a child is moved to a higher index, all children from the one just above the child index up to and including the target index are moved down by 1, and the child is assigned the target index value.
In general, if object a is above object b, the following code to moves a directly below b:
setChildIndex( a, getChildIndex( b ) );
Conversely, if object a is below object b, the preceding code moves a directly above b.
So far, we've always been moving around children that we've had a reference to. For example, the blue variable referenced the display object for the blue circle, and we were able to use this variable to change the index of the blue circle. What happens when you don't have a reference to the object you want to move, and the blue variable doesn't exist? The setChildIndex( ) method requires a reference to the object as its first parameter, so you'll need to get the reference somehow if it isn't available with a regular variable. The solution is to use the getChildAt( ) method.
The getChildAt( ) method takes a single argument, an index in the container's children list, and returns a reference to the display object located at that index. If the specified index isn't a valid index in the list, a RangeError is thrown.
The following example creates several circles of various colors and sizes and places them at various locations on the screen. Every time the mouse is pressed, the child at the very bottom is placed on top of all of the others:
package {
import flash.display.*;
import flash.events.*;
public class GetChildAtExample extends Sprite {
public function GetChildAtExample( ) {
// Define a list of colors to use
var color:Array = [ 0xFF0000, 0x990000, 0x660000, 0x00FF00,
0x009900, 0x006600, 0x0000FF, 0x000099,
0x000066, 0xCCCCCC ];
// Create 10 circles and line them up diagonally
for ( var i:int = 0; i < 10; i++ ) {
var circle:Shape = createCircle( color[i], 10 );
circle.x = i;
circle.y = i + 10; // the + 10 adds padding from the top
addChild( circle );
}
stage.addEventListener( MouseEvent.CLICK, updateDisplay );
}
// Move the circle at the bottom to the very top
public function updateDisplay( event:MouseEvent ):void {
// getChildAt(0) returns the display object on the
// very bottom, which then gets moved to the top
// by specifying index numChildren - 1 in setChildIndex
setChildIndex( getChildAt(0), numChildren - 1 );
}
// Helper function to create a circle shape with a given color
// and radius
public function createCircle( color:uint, radius:Number ):Shape {
var shape:Shape = new Shape( );
shape.graphics.beginFill( color );
shape.graphics.drawCircle( 0, 0, radius );
shape.graphics.endFill( );
return shape;
}
}
}
Recipes 6.1 and 6.2
You want to create a new type of DisplayObject.
DisplayObject
Create a new class that extends DisplayObject or one of its subclasses so it can be added into a display object container via addChild( ) or addChildAt( ).
addChild( )
addChildAt( )
Among the benefits of moving toward the display list model is the ease of creating new visual classes. In the past, it was possible to extend MovieClip to create custom visuals, but there always had to be a MovieClip symbol in the library linked to the ActionScript class to create an on-screen instance via attachMovie( ). Creating a custom visual could never be done entirely in ActionScript. With the display list model, the process has been simplified, allowing you to do everything in pure ActionScript code in a much more intuitive manner.
MovieClip
attachMovie( )
In the display list model, as discussed in the introduction of this chapter, there are many more display classes available besides just MovieClip. Before you create your custom visual, you need to decide which type it is going to be. If you're just creating a custom shape, you'll want to extend the Shape class. If you're creating a custom button, you'll probably want to extend SimpleButton. If you want to create a container to hold other display objects, Sprite is a good choice if you don't require the use of a timeline. If you need a timeline, you'll need to subclass MovieClip.
Shape
SimpleButton
Sprite
All of the available display object classes are tailored for specific purposes. It's best to decide what purpose your own visual class is going to serve, and then choose the appropriate parent class based on that. By choosing the parent class carefully you optimize size and resource overhead. For example, a simple Circle class doesn't need to subclass MovieClip because it doesn't need the timeline. The Shape class is the better choice in this case because it's the most lightweight option that appropriately fits the concept of a circle.
Circle
Once the base class has been decided, all you need to do is write the code for the class. Let's follow through with the circle example and create a new Circle class that extends the Shape display object. In a new ActionScript file named Circle.as, enter the following code:
Circle.as
package {
import flash.display.Shape;
/* The Circle class is a custom visual class */
public class Circle extends Shape {
// Local variables to store the circle properties
private var _color:uint;
private var _radius:Number;
/*
* Constructor: called when a Circle is created. The default
* color is black, and the default radius is 10.
*/
public function Circle( color:uint = 0x000000, radius:Number = 10 ) {
// Save the color and radius values
_color = color;
_radius = radius;
// When the circle is created, automatically draw it
draw( );
}
/*
* Draws the circle based on the color and radius values
*/
private function draw( ):void {
graphics.beginFill( _color );
graphics.drawCircle( 0, 0, _radius );
graphics.endFill( );
}
}
}
The preceding code defines a new Circle display object. When a Circle instance is created, you can specify both a color and a radius in the constructor. Methods from the Drawing API (discussed in Recipe 7.3) are used to create the body of the circle with the graphics property, which is inherited from the superclass Shape.
graphics
It is always a good idea to separate all drawing code into a separate draw( ) method. The constructor for Circle does not draw the circle directly, but it calls the draw( ) method to create the visual elements.
draw( )
All that is left to do is create new instances of our custom Circle class and add them to the display list with addChild( ) or addChildAt( ) so they appear on-screen. To create new instances of the class, use the new keyword. The following code example creates a few Circle instances and displays them on the screen:
new
package {
import flash.display.Sprite;
public class UsingCircleExample extends Sprite {
public function UsingCircleExample( ) {
// Create some circles with the Circle class and
// change their coordinates so they are staggered
// and aren't all located at (0,0).
var red:Circle = new Circle( 0xFF0000, 10 );
red.x = 10;
red.y = 20;
var green:Circle = new Circle( 0x00FF00, 10 );
green.x = 15;
green.y = 25;
var blue:Circle = new Circle( 0x0000FF, 10 );
blue.x = 20;
blue.y = 20;
// Add the circles to the display list
addChild( red );
addChild( green );
addChild( blue );
}
}
}
Recipes 6.1 and 7.3
You want to create an interactive button that enables a user to click and perform an action, such as submitting a form or calculating a total.
Create an instance of the SimpleButton class and create display objects for upState, downState, overState, and hitTestState. Alternatively, create a subclass of SimpleButton that describes your desired button behavior.
upState
downState
overState
hitTestState
Use the click event to invoke a method whenever the user presses the button.
click
The display list model provides an easy way to create buttons through the SimpleButton class. The SimpleButton class allows a user to interact with the display object using their mouse, and makes it easy for you to define that interaction through various button states. The possible button states, listed here, are available as properties of the SimpleButton class:
Pages: 1, 2, 3, 4
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Memory leaks are a class of bugs where memory is not released even after it is no longer needed. They are often explicit, and highly visible, which makes them a great candidate to begin learning debugging. Go is a language particularly well suited to identifying memory leaks because of its powerful toolchain, which ships with amazingly capable tools (pprof) which make pinpointing memory usage easy.
I’m hoping this post will illustrate how to visually identify memory, narrow it down to a specific process, correlate the process leak with work, and finally find the source of the memory leak using pprof. This post is contrived in order to allow for the simple identification of the root cause of the memory leak. The pprof overview is intentionally brief and aims to illustrate what it is capable of and isn’t an exhaustive overview of its features.
The service used to generate the data in this post is available here.
If memory grows unbounded and never reaches a steady state then there is probably a leak. The key here is that memory grows without ever reaching a steady state, and eventually causes problems through explicit crashes or by impacting system performance.
Memory leaks can happen for any number of reasons. There can be logical leaks where data-structures grow unbounded, leaks from complexities of poor object reference handling, or just for any number of other reasons. Regardless of the source, many memory leaks elicit a visually noticeable pattern: the “sawtooth”.
This blog post is focused on exploring how to identify and pinpoint root cause for a go memory leak. We’ll be focusing primarily on characteristics of memory leaks, how to identify them, and how to determine their root cause using go. Because of this our actual debug process will be relatively superficial and informal.
The goal of our analysis is to progressively narrow scope of the problem by whittling away possibilities until we have enough information to form and propose a hypothesis. After we have enough data and reasonable scope of the cause, we should form a hypothesis and try to invalidate it with data.
Each step will try to either pinpoint a cause of an issue or invalidate a non-cause. Along the way we’ll be forming a series of hypotheses, they will be necessarily general at first then progressively more specific. This is loosely based on the scientific method. Brandon Gregg does an amazing job of covering different methodologies for system investigation (primarily focused on performance).
Just to reiterate we’ll try to:
- Ask a question
- Form a Hypothesis
- Analyze the hypothesis
- Repeat until the root cause is found
How do we even know if there is a problem (ie memory leak)? Explicit errors are direct indicators of an issue. Common errors for memory leaks are: OOM errors or explicit system crashes.
Errors are the most explicit indicator of a problem. While user/application generated errors have the potential to generate false positives if the logic is off, an OOM error is the OS literally indicating something is using too much memory. For the error listed below this manifests as cgroup limits being reached and the container being killed.
dmesg
Question: Is the error a regular repeating issue?
Hypothesis: OOM errors are significant enough that they should rarely occur. There is a memory leak in one of the processes.
Prediction: Either the Process memory limit has been set too low and there was a uncharacteristic bump or there is a larger issue.
Test: Upon further inspection there are quite a few OOM errors suggesting this is a serious issue and not a one off. Check the system memory for historic view into memory usage.
The next step after identifying a potential problem is to get an idea of system wide memory usage. Memory Leaks frequently display a “sawtooth” pattern. The spikes correspond to the application running while the dips correspond to a service restart.
Sawtooth characterizes a memory leak especially corresponding with a service deploy. I’m using a test project to illustrate memory leaks but even a slow leak would look like saw tooth if the range is zoomed out far enough. With a smaller time range it would look like a gradual rise and then a drop off on process restart.
The graph above shows the example of a sawtooth memory growth. Memory continually grows without flatlining. This is a smoking gun for memory issues.
Question: Which process (or processes) is (are) responsible for the memory growth?
Test: Analyze per process memory. There could be information in the dmesg logs to indicate a process or class of processes that are the targets of OOM.
Once a memory leak is suspected the next step is to identify a process that is contributing, or causing the system memory growth. Having per process historical memory metrics is a crucial requirement (container based system resources are available through the a tool like cAdvisor). Go’s prometheus client provides per process memory metrics by default, which is where the graph below gets its data.
The below graph shows a process which is very similar to the system sawtooth memory leak graph above: continual growth until process restarts.
Memory is a critical resource and can be used to indicate abnormal resource usage or could be used as a dimension for scaling. Additionally, having memory stats help inform how to set container based (cgroups) memory limits. The specifics of the graph values above can be found metric code source. After the process has been identified it’s time to dig in and find out which specific part of the code is responsible for this memory growth.
Once again prometheus gives us detailed information into the go runtime, and what our process is doing. The chart shows that bytes are continually allocated to the heap until a restart. Each dip corresponds with the service process restart.
Question: Which part(s) of the application is(are) leaking memory?
Hypothesis: There’s a memory leak in a routine which is continually allocating memory to the heap (global variable or pointer, potentially visible through escape analysis)
Test: Correlate the memory usage with an event.
Establishing a correlation will help to partition the problem space by answering: is this happening online (in relation to transactions) or in the background?
One way to determine this could be to start the service and let it idle without applying any transactional load. Is the service leaking? If so it could be the framework or a shared library. Our example happens to have a strong correlation with transactional workload.
The above graph show the count of HTTP requests. These directly match the system memory growth and time and establish diving into HTTP request handling as a good place to start.
Question: Which part of the application are responsible for the heap allocations?
Hypothesis: There is an HTTP handler that is continually allocating memory to the heap.
Test: Periodically analyze heap allocations during program running in order to track memory growth.
In order to inspect how much memory is being allocated and the source of those allocations we’ll use pprof. pprof is an absolutely amazing tool and one of the main reasons that I personally use go. In order to use it we’ll have to first enable it, and then take some snapshots. If you’re already using http, enabling it is literally as easy as:
import _ "net/http/pprof"
Once pprof is enabled we’ll periodically take heap snapshots throughout the life of process memory growth. Taking a heap snapshot is just as trivial:
curl > heap.0.pprof
sleep 30
curl > heap.1.pprof
sleep 30
curl > heap.2.pprof
sleep 30
curl > heap.3.pprof
The goal is to get an idea of how memory is growing throughout the life of the program. Let’s inspect the most recent heap snapshot:
This is absolutely amazing. pprof defaults to
Type: inuse_space which displays all the objects that are currently in memory at the time of the snapshot. We can see here that
bytes.Repeat is directly responsible for 98.60% of all of our memory!!!
The line below the
bytes.Repeat entry shows:
1.28MB 0.31% 98.91% 410.25MB 98.91% main.(*RequestTracker).Track
This is really interesting, it shows that
Track itself has
1.28MB or
0.31% but is responsible for
98.91% of all in use memory!!!!!!!!!!!!! Further more we can see that http has even less memory in use but is responsible for even more than
Track (since
Track is called from it).
pprof exposes many ways to introspect and visualize memory (in use memory size, in use number of objects, allocated memory size, allocated memory objects), it allows listing the track method and showing how much each line is responsible for:
This directly pinpoints the culprit:
1.28MB 410.25MB 24: rt.requests = append(rt.requests, bytes.Repeat([]byte("a"), 10000))
pprof also allows visual generation of the textual information above:
(pprof) svg
Generating report in profile003.svg
This clearly shows the current objects occupying the process memory. Now that we have the culprit
Track we can verify that it is allocating memory to a global without ever cleaning it up, and fix the root issue.
Resolution: Memory was being continually allocated to a global variable on each HTTP request. | https://www.tefter.io/bookmarks/85685/readable | CC-MAIN-2019-43 | refinedweb | 1,558 | 55.34 |
Created on 2019-02-18 00:00 by rhettinger, last changed 2020-01-28 03:40 by rhettinger. This issue is now closed.
Attached is a class that I've found useful for doing practical statistics work with normal distributions. It provides a nice, high-level API that makes short-work of everyday statistical problems.
------ Examples --------
# Simple scaling and translation
temperature_february = NormalDist(5, 2.5) # Celsius
print(temperature_february * (9/5) + 32) # Fahrenheit
# Classic probability problems
#
# The mean score on a SAT exam is 1060 with a standard deviation of 195
# What percentage of students score between 1100 and 1200?
sat = NormalDist(1060, 195)
fraction = sat.cdf(1200) - sat.cdf(1100)
print(f'{fraction * 100 :.1f}% score between 1100 and 1200')
# Combination of normal distributions by summing variances
birth_weights = NormalDist.from_samples([2.5, 3.1, 2.1, 2.4, 2.7, 3.5])
drug_effects = NormalDist(0.4, 0.15)
print(birth_weights + drug_effects)
# Statistical calculation estimates using simulations
# Estimate the distribution of X * Y / Z
n = 100_000
X = NormalDist(350, 15).examples(n)
Y = NormalDist(47, 17).examples(n)
Z = NormalDist(62, 6).examples(n)
print(NormalDist.from_samples(x * y / z for x, y, z in zip(X, Y, Z)))
# Naive Bayesian Classifier
#])
prior_male = 0.5
prior_female = 0.5
posterior_male = prior_male * height_male.pdf(6) * weight_male.pdf(130) * foot_size_male.pdf(8)
posterior_female = prior_female * height_female.pdf(6) * weight_female.pdf(130) * foot_size_female.pdf(8)
print('Predict', 'male' if posterior_male > posterior_female else 'female')
I like this idea!
Should the "examples" method be re-named "samples"? That's the word used in the docstring, and it matches the from_samples method.
+1, This would be useful for quick analyses, avoiding the overhead of installing scipy and looking through its documentation.
Given that it's in the statistics namespace, I think the name can be simply ``Normal`` rather than ``NormalDist``. Also, instead of ``.from_examples`` consider naming the classmethod ``.fit``.
I'll work up a PR for this.
We can continue to tease-out the best method names. I've has success with "examples" and "from_samples" when developing this code in the classroom. Both names had the virtue of being easily understood and never being misunderstood.
Intellectually, the name fit() makes sense because we are using data to create best fit model parameters. So, technically this is probably the most accurate terminology. However, it doesn't match how I think about the problem though -- that is more along the lines of "use sampling data to make a random variable with a normal distribution". Another minor issue is that class methods are typically (but not always) recognizable by their from-prefix (e.g. dict.fromkeys, datetime.fromtimestamp, etc).
"NormalDist" seems more self explanatory to me that just "Normal". Also, the noun form seems "more complete" than a dangling adjective (reading "normal" immediately raises the question "normal what?"). FWIW, MS Excel also calls their variant NORM.DIST (formerly spelled without the dot).
Okay the PR is ready.
If you all are mostly comfortable with it, it would great to get this in for the second alpha so that people have a chance to work with it.
Thanks Raymond.
Apologies for commenting here instead of at the PR.
While I've been fighting with more intermittedly broken than usual
internet access, Github has stopped supporting my browser. I can't
upgrade the browser without upgrading the OS, and I can't upgrade the OS
without new hardware, and that will take money I don't have at the moment.
So the bottom line is that while I can read *part* of the diffs on
Github, that's about all I can do. I can't comment there, I can't fork,
I can't make push requests, half the pages don't load for me and the
other half don't work properly when they do load. I can't even do a git
clone.
So right now, the only thing I can do is comment on your extensive
documentation in statistics.rst. That's very nicely done.
The only thing that strikes me as problematic is the default value for
sigma, namely 0.0. The PDF for normal curve divides by sigma, so if
that's zero, things are undefined. So I think that sigma ought to be
strictly positive.
I also think it would be nice to default to the standard normal curve,
with mu=0.0 and sigma=1.0. That will make it easy to work with Z scores.
Thanks again for this class, and my apologies for my inability to
follow the preferred workflow.
I've made both suggested changes, "examples"->"samples" and set the defaults to the standard normal distribution.
To bypass Github, I've attached a diff to this tracker issue. Let me know what you think :-)
@steven.daprano Bit off topic but you can also append .patch in the PR URL to generate patch file with all the commits made in the PR up to latest commit and .diff provides the current diff against master. They are plain text and can be downloaded through wget and viewed with an editor in case if it helps.
Thanks for all the positive feedback. If there are no objections, I would like to push this so it will be in the second alpha release so that it can get exercised. We can still make adjustments afterwards.
New changeset 11c79531655a4aa3f82c20ff562ac571f40040cc by Raymond Hettinger in branch 'master':
bpo-36018: Add the NormalDist class to the statistics module (GH-11973)
Okay, it's in for the second alpha. Please continue to make API or implementation suggestions. Nothing is set in stone.
There is an inconsistency worth paying attention to in the choice of names of the input parameters.
Currently in the statistics module, pvariance() accepts a parameter named "mu" and pstdev() and variance() each accept a parameter named "xbar". The docs describe both "mu" and "xbar" as "it should be the mean of data". I suggest it is worth rationalizing the names used within the statistics module for consistency before reusing "mu" or "xbar" or anything else in NormalDist.
Using the names of mathematical symbols that are commonly used to represent a concept is potentially confusing because those symbols are not always *universally* used. For example, students are often introduced to new concepts in introductory mathematics texts where concepts such as "mean" appear in formulas and equations not as "mu" but as "xbar" or simply "m" or other simple (and hopefully "friendly") names/symbols. As a mathematician, if I am told a variable is named, "mu", I still feel the need to ask what it represents. Sure, I can try guessing based upon context but I will usually have more than one guess that I could make.
Rather than continue down a path of using various mathematical-symbols-written-out-in-English-spelling, one alternative would be to use less ambiguous, more informative variable names such as "mean". It might be worth considering a change to the parameter names of "mu" and "sigma" in NormalDist to names like "mean" and "stddev", respectively. Or perhaps "mean" and "standard_deviation". Or perhaps "mean" and "variance" would be easier still (recognizing that variance can be readily computed from standard deviation in this particular context). In terms of consistency with other packages that users are likely to also use, scipy.stats functions/objects commonly refer to these concepts as "mean" and "var".
I like the idea of making NormalDist readily approachable for students as well as those more familiar with these concepts. The offerings in scipy.stats are excellent but they are not always the most approachable things for new students of statistics.
Karthikeyan: thanks for the hint about Github.
Raymond: thanks for the diff. Some comments:
Why use object.__setattr__(self, 'mu', mu) instead of self.mu = mu in the __init__ method?
Should __pos__ return a copy rather than the instance itself?
The rest looks good to me, and I look forward to using it.
Davin: the chice of using mu versus xbar was deliberate, as they represent different quantities: the population mean versus a sample mean. But reading over the docs with fresh eyes, I can now see that the distinction is not as clear as I intended.
I think that changing the names now would be a breaking change, but even if it wasn't, I don't want to change the names. The distinction between population parameters (mu) and sample statistics (xbar) is important and I think the function parameters should reflect that.
As for the new NormalDist class, we aren't limited by backwards compatibility, but I would still argue for the current names mu and sigma. As well as matching the population parameters of the distribution, they also matches the names used in calculators such as the TI Nspire and Casio Classpad (two very popular CAS calculators used by secondary school students).
See #36099. If you would like to suggest some doc changes, please feel free to do so.
Steven: Your point about population versus sample makes sense and your point that altering their names would be a breaking change is especially important. I think that pretty well puts an end to my suggestion of alternative names and says the current pattern should be kept with NormalDist.
I particularly like the idea of using the TI Nspire and Casio Classpad to guide or help confirm what symbols might be recognizable to secondary students or 1st year university students.
Raymond: As an idea for examples demonstrating the code, what about an example where a plot of pdf is created, possibly for comparison with cdf? This would require something like matplotlib but would help to visually communicate the concepts of pdf, perhaps with different sigma values?
> Why use object.__setattr__(self, 'mu', mu) instead of
> self.mu = mu in the __init__ method?
The idea was the instances should be immutable and hashable, but this added unnecessary complexity, so I took this out prior to the check in.
> Should __pos__ return a copy rather than the instance itself?
Yes. I'll fix that straight-way.
^ The chice of using mu versus xbar was deliberate
I concur with that choice and also prefer to stick with mu and sigma:
1) It's too late to change it elsewhere in statistics and the random modules. 2) Having attribute names the same as function names in the same module is confusing. 3) I had already user tested this API in some Python courses. 4) The variable names match the various external sources I've linked to in the docs. 5) Python historically hasn't shied from greek letter names (math: pi tau gamma random: alpha, better, lambd, mu, sigma).
Steven, Davin, Michael: Thanks for the encouragement and taking the time to review this code.
New changeset 79fbcc597dfd039d3261fffcb519b5ec5a18df9d by Miss Islington (bot) (Raymond Hettinger) in branch 'master':
bpo-36018: Make __pos__ return a distinct instance of NormDist (GH-12009)
New changeset 9e456bc70e7bc9ee9726d356d7167457e585fd4c by Miss Islington (bot) (Raymond Hettinger) in branch 'master':
bpo-36018: Add properties for mean and stdev (GH-12022)
New changeset ef17fdbc1c274dc84c2f611c40449ab84824607e by Miss Islington (bot) (Raymond Hettinger) in branch 'master':
bpo-36018: Add special value tests and make minor tweaks to the docs (GH-12096)
New changeset 9add4b3317629933d88cf206a24b15e922fa8941 by Miss Islington (bot) (Raymond Hettinger) in branch 'master':
bpo-36018: Add documentation link to "random variable" (GH-12114)
I've done some spot checks of NormDist.pdf and .cdf and compared the results to those returned by my TI Nspire calculator.
So far, the PDF has matched that of the Nspire to 12 decimal places (the limit the calculator will show), but the CDF differs on or about the 8th decimal place:
py> x = statistics.NormalDist(2, 1.3)
py> x.cdf(5.374)
0.9952757439207682
# Nspire normCdf(-∞, 5.372, 2, 1.3) returns 0.995275710979
# difference of 3.294176820212158e-08
py> x.cdf(-0.23)
0.04313736707891003
# Nspire normCdf(-∞, -0.23, 2, 1.3) returns 0.043137332077
# difference of 3.500191003008579e-08
Wolfram Alpha doesn't help me decide which is correct, as it doesn't show enough decimal places.[+NormalDistribution[2,+1.3],+5.374+][+NormalDistribution[2,+1.3],+-0.23+]
Do we care about this difference? Should I raise a new ticket for it?
According to GP/Pari, the correctly value for the first result, to the first few dozen places, is:
0.995275743920768157605659214368609706759611629000344854339231928536087783251913252354...
I'm assuming you meant 5.374 rather than 5.372 in the first Nspire result.
Below is the full transcript from Pari/GP: note that I converted the float inputs to exact Decimal equivalents, assuming IEEE 754 binary64. Summary: both Python results look fine; it's Nspire that's inaccurate here.
mirzakhani:~ mdickinson$ /opt/local/bin/gp
GP/PARI CALCULATOR Version 2.11.1 (released)
i386 running darwin (x86-64/GMP-6.1.2 kernel) 64-bit version
compiled: Jan 24 2019, Apple LLVM version 10.0.0 (clang-1000.11.45.5)
threading engine: single
(readline v8.0 enabled, extended help enabled)
PARI/GP is free software, covered by the GNU General Public License, and comes WITHOUT ANY WARRANTY WHATSOEVER.
Type ? for help, \q to quit.
Type ?17 for how to get moral (and possibly technical) support.
parisize = 8000000, primelimit = 500000
? \p 200
realprecision = 211 significant digits (200 digits displayed)
? ncdf(x, mu, sig) = (2 - erfc((x - mu) / sig / sqrt(2))) / 2
%1 = (x,mu,sig)->(2-erfc((x-mu)/sig/sqrt(2)))/2
? ncdf(5.37399999999999966604491419275291264057159423828125, 2, 1.3000000000000000444089209850062616169452667236328125)
%2 = 0.99527574392076815760565921436860970675961162900034485433923192853608778325191325235412640687571628164064779657215907190523884572141701976336760387216713270956350229484865180142256611330976179584951493
? ncdf(-0.2300000000000000099920072216264088638126850128173828125, 2, 1.3000000000000000444089209850062616169452667236328125)
%3 = 0.043137367078910025352120502108682523151629166877357644882244088336773338416883044522024586619860574718679715351558322591944140762629090301623352497457372937783778706411712862062109829239761761597057063
> I'm assuming you meant 5.374 rather than 5.372 in the first Nspire result.
Yes, that was a typo, sorry.
Thanks for checking into the results.
New changeset fb8c7d53326d137785ca311bfc48c8284da46770 by Raymond Hettinger in branch 'master':
bpo-36018: Make "seed" into a keyword only argument (GH-12921)
New changeset b0a2c0fa83f9b79616ccf451687096542de1e6f8 by Raymond Hettinger in branch 'master':
bpo-36018: Test idempotence. Test two methods against one-another. (GH-13021)
New changeset 671d782f8dc52942dc8c48a513bf24ff8465b112 by Raymond Hettinger in branch 'master':
bpo-36018: Update example to show mean and stdev (GH-13047)
I have a query about the documentation:
The default *method* is "exclusive" and is used for data sampled
from a population that can have more extreme values than found
in the samples. ...
Setting the *method* to "inclusive" is used for describing
population data or for samples that include the extreme points.
In all my reading about quantile calculation methods, this is the first time I've come across this recommendation. Do you have a source for it or a justification?
Thanks.
New changeset 4db25d5c39e369f4b55eab52dc8f87f390233892 by Raymond Hettinger in branch 'master':
bpo-36018: Address more reviewer feedback (GH-15733)
New changeset cc1bdf91d53b1a4751be84ef607e24e69a327a9b by Raymond Hettinger in branch '3.8':
[3.8] bpo-36018: Address more reviewer feedback (GH-15733) (GH-15734)
New changeset 10355ed7f132ed10f1e0d8bd64ccb744b86b1cce by Raymond Hettinger in branch 'master':
bpo-36018: Add another example for NormalDist() (#18191)
New changeset eebcff8c071b38b53bd429892524ba8518cbeb98 by Raymond Hettinger (Miss Islington (bot)) in branch '3.8':
bpo-36018: Add another example for NormalDist() (GH-18191) (GH-18192)
New changeset 01bf2196d842fc20667c5336e0a7a77eb4fdc25c by Raymond Hettinger in branch 'master':
bpo-36018: Minor fixes to the NormalDist() examples and recipes. (GH-18226)
New changeset 41f4dc3bcf30cb8362a062a26818311c704ea89f by Raymond Hettinger (Miss Islington (bot)) in branch '3.8':
bpo-36018: Minor fixes to the NormalDist() examples and recipes. (GH-18226) (GH-18227) | https://bugs.python.org/issue36018 | CC-MAIN-2020-40 | refinedweb | 2,544 | 65.93 |
#!/usr/bin/perl
use strict; use warnings;
my %data;
my $InputFile = "/the/file/in/question";
open INPUT, "< $InputFile";
foreach $line (<INPUT>) {
my @array = split /:/, $line;
$data{$array[1]} = $array[2];
}
contents of /the/file/in/question:
foo:FOO
bar:BAR
baz:BAZ
qux:QUX
[download]
Check out the Config:: namespace on CPAN. There are quite a few modules that exist for this purpose - pick one that uses the most convenient format for you. I used Config::General a few times, but Config::IniFiles and Config::Simple also look good. I selected Config::General because it allowed me to insert comments and blank lines for whitespace, and also use a block structure to nest levels of config parameters (thereby giving me the ability to construct a HoH or other complex data structure). Config::IniFiles reads INI style files and Config::General reads a INI-like files.
How about you code your hash as %config=(foo=>'foo',bar=>'bar') etc., and just require it, using "our"? That way you don't have to read the file and convert it to Perl, it's Perl already.
($_='kkvvttuu bbooppuuiiffss qqffssmm iibbddllffss')
=~y~b-v~a-z~s; print
The following code is a slightly more robust version. Note in particular that the result of the open is checked and a three parameter open is used. Note too that the line gets chomped and the format checked (at least a little) and bad input lines rejected. Even then a lot more checking ought to be done (duplicate keys anyone?).
I'd recommend however that you look at some of the configuration options bobf mentioned.
#!/usr/bin/perl
use strict; use warnings;
my %data;
#open INPUT, '<', $InputFile" or die "Failed to open $InputFile: $!";
while (<DATA>) {
chomp;
next if ! /(\w+)\s*:\s*(\w+)/;
$data{$1} = $2;
}
print "$_ => $data{$_}\n" for keys %data;
__DATA__
foo:FOO
bar:BAR
bogus data line - comment maybe?
baz:BAZ
qux:QUX
[download]
Prints:
bar => BAR
baz => BAZ
qux => QUX
foo => FOO
[download]
my @array = split /:/, $line;
[download]
errormessage:"couldn't open file: check permissions"
[download]
module:LWP::Simple
[download]
my @array = split (/:/, $line, 2);
[download]
Any improvement on the OP would be a matter of robustness (along the lines suggested by GrandFather), or of "standardizing" on a solution that is already available (i.e. using a CPAN module), or merely of style or perceived maintainability (e.g. using fewer lines of Perl code and/or adding commentary/POD to describe the expected input file format, etc).
All of those are possible, but none of them have any impact that I could imagine on "making better use of the computer's resources". I'm not really sure what you mean by that, but if you mean "make the process more efficient", I don't think any change of the OP code would have noticeable impact -- what you've posted is close enough to being as efficient as possible.
If there is some other aspect to "use of the computer's resources" that you're thinking of, that might make an interesting discussion.
BTW, I think, given the sample file contents, your use of $date{$array[1]} = $array[2] would be wrong; the indexes should be 0 and 1 instead. Or better yet, something like this:
open INPUT, "<", $InputFile;
my %data = map { chomp; split( /:/, $_, 2 ) } <INPUT>;
close INPUT;
[download]
That's really minimalist -- maybe a bit too much so for some, but it's really a matter of taste and how much you can trust your data files to be as expected.
(Updated the split so that it returns at most two elements per line from the input file; this is still vulnerable to serious trouble if the file contains any sort of line that lacks a colon.)
"make better use of the computer's resources" is a pretty meaningless phrase - does it mean make the process as inefficient as possible so more computer resources are used? If it means "make this process execute in minimum time", then don't worry about it, even if the configuration file is megabytes big the time to slurp the file and process it (you did realise that the foreach slurps the file?) the time to process it is likely to be negligable. The time wasted diagnosing and fixing input errors due to a complete lack of validation on the other hand could cost a heap of time better spent drinking beer.
Get someone else to do the work for you - use a module, then take the rest of the afternoon off for a beer.
I'd say take that advice with a grain of salt... For things that are basically pretty simple -- easy to set up and easy to validate, like the OP's case -- it can be quicker and more reliable to roll your own. A module posted by someone else may have been written to do a slightly different task, and finding that out, and figuring out whether/how it can be shoe-horned into your particular task, might end up being more work with a less satisfying outcome.
But for the harder things that make you scratch your head and say "I'm not sure how to solve this", definitely go to CPAN and look for help. Even if there is no single module that does exactly what you need, you're likely to learn about how to break the problem down into manageable chunks, and/or find useful references, and so on.
Many things look simple, few things are simple. Take OP's "simple requirements" for example: "I'd like my program to read in some data at start, rather than hard-code it into the program". Ok, OP is talking about storing some configuration information and the sample data given indicates simple key:value data.
But how simple is that? The code given breaks in all sorts of ways - empty lines, lines without a colon, values containing colons, duplicate keys, nasty configuration file names (two param open rather than three), missing or otherwise unreadable configuration file and very likely other things as yet unthought of.
Ok, OP spends a little time up front trawling through CPAN (with a little guidance) and comes up with a tool kit for solving the problem today and, well golly, solving the problem again in a different context tomorrow. Sounds like well spent time to me, and there is still time left this afternoon for a beer.
Sure, at some point you have to write some code to solve your own specific problem, but the more glue and the less new code the more likely it is that you don't have to deal with all the edge cases and stuff you've not thought of. In this case half an hour research and five minutes coding is likely to save several hours down the track bodging up the holes in the first implementation - and they are likely to be hours with people breathing down your neck as you sort out problems with a live system. Saving those sort of hours is worth several up front hours any day!
ok, here's one:
#!/usr/bin/perl
use strict; use warnings;
use Tie::File;
tie my @array, 'Tie::File', "/the/file/in/question" or die "can't tie
+file";
my %data = map { split /:/; $_[0] => $_[1] } @array;
| http://www.perlmonks.org/?node_id=567118 | CC-MAIN-2014-23 | refinedweb | 1,224 | 63.93 |
Haskell/Preliminaries
Note
This page is not part of the main Haskell Wikibook. It overlaps with the primary introductory material but isn't as polished. It has some assumptions about imperative programming background. Because learning works best when experiencing things from many angles, reading this as a supplement to reinforce basic concepts may be more effective than simply re-reading other chapters. Still, this page is basically extraneous.
(All the examples in this chapter can be typed into a Haskell source file and evaluated by loading that file into GHC or Hugs.)
Variables that Don't and Functions that Do[edit]
One of the main ideas behind functional programming is referential transparency — the idea that a function and a parameter can always be replaced by the function's return value. For example, the expression
2+2 is interchangeable with the value
4. This facilitates a number of important concepts in Haskell including lazy evaluation and the ability to reorder expressions to execute them concurrently.
Unlike imperative languages such as C++, Haskell does not treat variables as memory locations. In Haskell, variables are bound to a particular value within whatever scope the variable exists. They are more like constants than variables in most cases. The concept of variable in math is actually the same as that in Haskell. In the following example, x is bound to 3; it can't possibly change.
x = 6 / 2
This may seem extremely weird if you are used to imperative programming. How can we write useful programs without actual variables? Well, Haskell's function parameters can serve the same purpose: once you pass a value to a function, that function can use it wherever it wants. To change the value of a parameter, just call the function again. So the answer is recursion.
Here are some Haskell declarations:
x = 3 y = x * 2
As you've probably guessed, by the end y = 6. Looks quite a bit like math doesn't it?
However, unlike a conventional language, the order doesn't matter. Haskell doesn't have a notion of "flow of control", so it doesn't start executing at the top and working its way down. So this means exactly the same as the above:
y = x * 2 x = 3
You can of course also define functions. This is where life starts to get a little more interesting.
double x = x * 2 z = double y
This says that the function "double" takes an argument "x" and the result is equal to twice x. Note the lack of parentheses: in Haskell "double y" means apply the function "double" to the value "y". Parentheses are only needed to determine evaluation priority, just like in math.
So far, this lets you do about as much as a spreadsheet would: variables in this world look a lot like cells in a spreadsheet, and you can set up functions that can "do" things with these variables.
Now look at this"; function application (as this process is known) has higher precedence than any regular operator like addition or exponentiation.
This shows how you do "loops" in Haskell. You don't have a loop counter like "n = n - 1" because that would change the value of "n". So instead you use recursion.
Unlike our earlier example, the order of the two recursive declarations is important. Haskell matches function calls starting at the top and picking the first one that matches. So, if the two declarations were reversed then the compiler would conclude that factorial 0 equals 0 * factorial -1, and so on to infinity. Not what we want.
Tail Recursion[edit]
In case you're wondering about stack space, won't all this recursion eventually overflow the stack? The answer is that Haskell supports tail recursion (every decent functional language's compiler does). Tail recursion is the idea that if the last thing a function does is call itself, the interpreter may as well not bother putting things on the stack, and just "jump" to the beginning using the function's new arguments. In this case, though, the factorial function has to call itself with "(n-1)" and then multiply the result by "n".
So our factorial function isn't tail recursive (because it has to multiply result by n after calling itself). This one is, though:
factorial n = factorial' n 1 factorial' 0 a = a factorial' n a = factorial' (n-1) (n*a)
Here are two functions. They have almost the same name, except the second has a tick mark. (Haskell identifiers are allowed to contain tick marks.) By convention, ticks are used to indicate a variation on something.
Here, the
factorial' function takes an argument
n as before, but it also takes a second "accumulator" argument. Because calling itself is the last thing it does, the tail recursion rule applies: nothing gets put on the stack; the
factorial' function simply executes as if its arguments had always been n-1 and n*a.
(Note the similarity between this parameter and a variable. Parameters are, in a way, the variables of Haskell.)
A programmer using this function doesn't want to worry about the accumulator argument, so the plain factorial function is used to "wrap up"
factorial' by supplying the initial value of the accumulator.
Exercises[edit]
Type the factorial function into a Haskell source file and load it into your favourite Haskell environment.
What is factorial 5?
What about factorial 1000? Is that what you expected?
What about factorial -1?
Types[edit]
You may have noticed something interesting above. None of your parameters or anything was declared to be of any particular type.
That's because Haskell uses a type system to infer the type of every expression, the Hindley-Milner type system. But Haskell is not weakly typed; it's strongly typed like C++, so you can't get runtime type errors. Any program that compiles is guaranteed to not have type errors in it. Another advantage is that when you make a mistake, it will almost always change the type of the resulting expression, and it'll be flagged by the compiler as such. (Your program simply won't compile.) You may still want to write types for your functions for a few reasons.
- It helps to document your program.
- When you get something wrong, the type declarations help to narrow down the source of the problem.
- Most importantly, it will help you learn. Understanding the type system is fundamental to getting a good grasp on Haskell.
The syntax for type declarations looks like this:
x :: Integer factorial :: Integer -> Integer
You should read the first declaration as saying "x is of type Integer" or "x has type Integer". The '::' tells us we're referring to a type. You'll note factorial looks a bit different. It's a function that takes an Integer and returns an Integer. Later on you'll see why we use -> in describing the types, and why describing the type of a function is so close syntactically to describing the type of a variable.
One final note: identifiers that start with uppercase letters are reserved for types, modules and data constructors (somewhat like namespaces, don't worry about these for now). If you try to declare a variable or function that begins with an uppercase letter, the compiler will complain. | https://en.wikibooks.org/wiki/Haskell/Preliminaries | CC-MAIN-2017-04 | refinedweb | 1,216 | 64 |
Why React?
Nowadays developers know that Javascript’s world is constantly changing. Every day a new library, framework or even an update to the language itself appears and we must be aware of all these as developers. Today I’m going to talk about one of many features of React JS: Loops in JSX but maybe you are wondering “What is this React JS thing?”.
React JS or just React is a Javascript library for building user interfaces, made by Facebook, they say React have many virtues, for example:
- Is declarative: That means interactive UI design is no longer a pain in the neck, and each component follows a state that responds to data changes.
Component-Based: You can build encapsulated components that are only aware of their own state, and by composing many of them, you can build more complex components.
Learn once, write anywhere: It’s technology stack agnostic, also you can build native and mobile apps with it.
New player: JSX
JSX describes itself as a statically-typed, object-oriented programming language designed to run on modern web browsers and most of the code written for React is using it. For example we have our first React component written using JSX:
class HelloMessage extends React.Component { render() { return <div>Hello {this.props.name}</div> ; } } ReactDOM.render(, mountNode);
I bet you noticed the XML-ish syntax here. Well, since React JSX transforms from an XML-like syntax into native JavaScript, we might need a compiler like Babel to get an output like this:
class HelloMessage extends React.Component { render() { return React.createElement( "div", null, "Hello ", this.props.name ); } } ReactDOM.render(React.createElement(HelloMessage, { name: "Jane" }), mountNode);
Also using JSX is optional, but our friends at Facebook recommend to use it because “It is a concise and familiar syntax for defining tree structures with attributes.”
Example: Creating a simple comments list
All right, now it’s time to throw some code here. First, we need to visualize our application in a React way, what does that mean? It means that as we said before, React allows us to define encapsulated components, so for this example, we only have a simple list of comments, and a quick way to define it would be like this:
Component: Single Comment or simply called Comment
Function: Display the comments parts, and this time we are keeping everything simple and showing three parts on each comment:
- Unique identifier (this could be an index or an id given by a database).
- The comment’s author.
- The content, which is whatever the author wrote.
Since we are isolating each component, it is unknown for this comment component where all the other comments are going to be displayed or the number of comments in the application. It just focuses on its function: which is displaying a single comment, with its properties
import React from 'react'; class Comment extends React.Component { constructor() { super(); } render() { return ( <div>#{this.props.id}</div> By: {this.props.author} {this.props.content} ); } } export default Comment;
As you can see, this component expects three properties:
– ID
– Author
– Content
And the way we access them is through the
props property, however, “where they come from?” – It is not important.
Component: List of Components or more simply called CommentList
Function: Display each one of the comments in a container, it should be responsible of the way the comments are displayed, grid, layout or whatever we desire.
This component will contain all the comments inside, in other words, a list of Comments
This is where the Loop comes to the game: in order to display the whole set of comments we first need to fetch them all from a source (Webservice, file, etc.) and then iterate the list and for each one of the comments, the CommentList should send the data for a single comment to a Comment component. Let’s see what I’m talking about.
/** Mock data, this could be fetched from a webservice **/ [ { "id":1, "author":"Esteban Cortes", "content":"Cool! This is so awesome!" }, { "id":2, "author":"Edwin Cruz", "content":"Let's get some wings and caguas to celebrate" }, { "id":3, "author":"Ernesto Alcaraz", "content":"But I forgot my crocs" }, { "id":4, "author":"Mario Gomez", "content":"Let them go!" } ]
// CommentList.js import React from 'react'; import Comment from './Comment'; // Notice we are importing the Component component class CommentList extends React.Component { constructor() { super(); /** When this component is instanced, it pass the reference of this into the getComments method **/ this.getComments = this.getComments.bind(this); } getComments() { this.comments = require('../mocks/comments'); } render() { this.getComments(); // We fetch the comments before the render return ( <div> <h2>MagmaComments</h2> {this.comments.map( (comment) => { return ( ); })} </div> ); } } export default CommentList;
The key code inside this is:
{this.comments.map( (comment) => { return ( ); })}
Getting everything together
Let’s dive into this:
- When the CommentList is called, it fetches the list of comments from a source, in this case from a mock, which is returned as an array of objects.
- In Javascript, the Array object has a method called
map, which iterates over each element inside an array and returns a brand new array after executing the callback in each one of the elements, so it’s called inside the JSX code inside the render method.
- In JSX we can use expressions inside the
{}characters, everything inside those will be evaluated and injected into the JSX template. So for each element inside the comments array, we create a Comment component and return it, creating a new array of Comment components.
- We share properties between objects using attributes, each attribute we set can be accessed using the
propsobject inside each component.
Then we access it:
render() { return ( <div> #{this.props.id}By: {this.props.author} {this.props.content} </div> ); }
Note: When creating multiple components React will require that each one has a unique id, we set this property using the
key attribute.
Finally we get something like this:
| http://blog.magmalabs.io/2016/10/18/reactjs-loops-in-jsx.html | CC-MAIN-2020-05 | refinedweb | 984 | 51.78 |
When i run
kubectl apply -f hpademo.yml
and then
kubectl get hpa –namespace acg-ns
i just get back
acg-hpa Deployment/acg-web /50% 1 10 1 18m
(as you can see i let it run for a long time)
I also noticed that
kubectl get nodes
just returned
NAME STATUS ROLES AGE VERSION
docker-desktop Ready control-plane,master 5h20m v1.21.1
there was no node for the pod.
so i never see a %cpu other than unknown and never see any replicas created when i run the stresser.
1 Answers
Docker desktop is a single node cluster, and the hpa is supposed to launch more nodes, so I think it only works with cloud clusters. | https://acloudguru.com/forums/kubernetes-deep-dive/the-hpa-demo-yml-file-is-not-creating-a-node | CC-MAIN-2021-49 | refinedweb | 120 | 74.53 |
I have been working on a Ruby on Rails project which has greedy asset
precompile regex (which is desirable in my case, because I'm not including
):
# in config/application.rb# this excludes all files
which start with an '_' character (sass)config.assets.precompile
<< /(?<!rails_admin)(^[^_/]|/[^_])([^/])*.s?css$/
In this same project, I a
I would like to know if there is a way to lock (prevent) an application
from starting.
And i also would like to know if there is a way
to prevent a service(application) from starting at boot of the
device
...i would like to know because i would like to create
an anti-malware app.
My code tries to extract the range even if the line starts with a * Here
is my code:
while (<FILE1>) { $_ =~ s/^s+//;
#remove leading spaces $_ =~ s/s+$//; #remove trailing spaces if (/IF/ .. /END-IF/) { if($_ =~ m/END-IF/) {
$flag = 1; } print FINAL "$_";
if ($flag == 1) {
I've been using the repeater control in asp.net for awhile now..and
every now and then i keep forgetting to add the '#' inside the < %#
DataBinder.Eval(Container.DataItem, "NAME") % >
so i was
wondering what does it mean ?
My C# MDI Application starts in program.cs, which looks like
this:
namespace APRSTW{ static class
Program { public static Form parentForm;
/// <summary> /// The main entry point for the
application. /// </summary> [STAThread] static void Main() {
I updated from elasticsearch 0.17.6 to 0.19.2 and everything seems fine,
but I cannot connect to port 9200 locally.
When I watch the es
log I see that it's cycling furiously through all node names without
actually connecting to any of them.
The logs look like this -
I have a MacBook Pro 13" (MD101) with Mountain Lion 10.8.2
installed.
I downloaded the ADT Bundle pack and when I tried to
run the "Hello World" program on an AVD, I found that the AVD won't load.
It just shows a blank screen in the skin window.
When I run
"adb logcat', it says "Waiting for device"When I run "adb shell", it
says "Device offline"When I run "ad
I can get the progress dialog to stop, but the TabbedView activity never
starts, just goes to a black screen. Any ideas?
class
DownloadWebPageTask extends AsyncTask<String, Void, String> {
private final ProgressDialog dialog = new
ProgressDialog(MainScreen.this); @Override protected
void onPreExecute() { dialog.setMessage | http://bighow.org/tags/Starting/1 | CC-MAIN-2017-47 | refinedweb | 403 | 63.19 |
This action might not be possible to undo. Are you sure you want to continue?
DEVELOPING STATIC WEBSITES
What we’ll cover in this chapter:
Site editing tools World Wide Web Consortium (W3C) XHTML Cascading Style Sheets JavaScript Templates and Library Items XHTML time
39
CHAPTER 3. In all of these scenarios, one set of users was guaranteed to come off worst of all: Mac users. I’d be willing to put all my Macs up as a stake and bet that nine out of ten designers who designed websites on a PC have never tested their work on a Mac. The sad fact is that the majority don’t seem to care. I, for one, always thought about both platforms, as I used both of them at work and home. I would often design a site at work using Windows as the test platform and the site would work just fine. Testing it at home on my Mac though was an entirely different story: There would be large gaps in the layout, misbehaving tables, and general funk occurring all over the place. A designer could make far too many mistakes and simply get away with it. “Missed some closing tags off? Not a problem, I’ll just render the page the way I think it should be!” said Internet Explorer, while Netscape would crank up the king of all tags (<blink>), and actually flash where the error was! Of course, a lot of this bad coding was done by hand and usually to impossible deadlines for evil tyrant dictators (a.k.a. The Client). No syntax coloring + no visual guides = No Fun, and a greater risk of mistakes being made. I regularly used to code 90K HTML pages in Notepad that were full of nested tables. Eating in a restaurant one night after work, a waiter said to me: “I’m afraid there’s a problem with your table, sir.” “Check both the <tr> and <td> tags are closed properly,” I said, without even thinking. The look of confusion on the guy’s face was classic, until I realized and explained. That’s when I knew things really had to be done a different way. I managed to convince my boss to let us use Dreamweaver, and deadlines were never as scary from that day forth. I started in this business by hand coding, and I’m still glad I did. I know how things work, and why. I can fix bits of code if I need to and make tiny adjustments without having to start up Dreamweaver. Hand coding is very useful, but let’s be realistic here. If a client is banging on the door, wondering where their overdue 200-page site is, it’s time for action.
Overview
For years, people have talked about making websites for a living or as a hobby. More often than not, it was very difficult to get these sites to work the same in the main two browsers (PC Internet Explorer and PC Netscape 4+). You would code some elements that would look fine in one browser, but in another it would look awful. Compromise was often the order of the day. The choices were Use a detection script to send the user to a browser-specific version of the site (a very popular technique that’s still in use today). Design purely for PC Internet Explorer, which is currently the most popular browser in use (a bad technique, as it blocks some users, creating a bad impression). Get the site looking “near enough” as you want it to in the main two browsers, and ignore all the other users. (Far too common. I’m sure some designers never knew there was life outside IE and Netscape.) Code properly! (More about this later.)
40
DEVELOPING STATIC WEBSITES This action comes in the form of What You See Is What You Get (WYSIWYG) web page editing software such as Dreamweaver or GoLive (these applications are discussed in detail in Chapter 2). Used in conjunction with their graphics counterparts (Fireworks and ImageReady, respectively, in the case of the Dreamweaver and GoLive), today’s developer can save a serious amount of time. In fact, you would have to be quite mad to shun tools like these. How about the ability to lay out your page, slice up the graphics, create image maps, add rollovers, add <alt> tags and status bar messages, then import all that into your HTML document with a single click? Does that do anything for you? It certainly makes things easier for me!
2. Fill in the User Info. 3. Type in the web address (see Figure 3-2). Let’s stay
local and use (I’ve left the example from the previous chapter in the root of my web server, to replace the default Apache index page).
Site editing tools
In Chapter 2, you looked at the heavyweight editors, Dreamweaver and GoLive. With full-on site management, visual CSS support, and everything else they have to offer, you can quite easily get lost in options if all you need to do is maintain an existing site. You might only want to make a few text changes. Or, maybe you’ve left users in charge of their own updates, and a program like Dreamweaver would be far too much for them to handle. Time to look at site editing tools!
Figure 3-2. Type in the local IP address for your web server.
4. Select Local/Network on the Connection Info page,
as shown in Figure 3-3.
Macromedia Contribute
Contribute is an interesting application that can be set up in a number of ways, depending on your requirements and users’ needs. First, imagine that you are the administrator of a site with total control and you’re giving your client only very basic rights, such as allowing the client to just edit text, change certain images, and so on. In a nutshell, you can allow your client to edit only certain regions of the site.
1. Click the Create Connection button, shown in
Figure 3-1, to invoke the Connection Assistant. Note the option to connect to your iDisk (which you can use, if you have a full .Mac account).
Figure 3-3. Choose the Local/Network option from the menu.
Figure 3-1. Click the button to start.
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CHAPTER 3
5. Still on the Connection Info page, browse to the
network path of your web server: MacintoshHD/ Library/WebServer/Documents (see Figure 3-4).
Figure 3-6. Click Yes, although you can do this part later too.
Administering users in Contribute
I’m going to set up Jake, my Technical Reviewer, as a user, and then have Contribute create a key to e-mail to him.
1. Click the Send Connection Key button, shown in
Figure 3-7.
Figure 3-4. Use the Browse button to select the path to the web server’s documents folder.
6. You want to administer the site, so click the Yes, I
want to be the administrator button and then enter the password you want to use (see Figure 3-5).
Figure 3-7. Hit the Send Connection Key button to start the export process.
2. Select the No radio button to customize the connection settings for others as shown in Figure 3-8 and click Continue.
Figure 3-5. Insert a password—make sure it’s one you’ll remember.
7. The Summary page should tell you that you’ve
been successful, so click Finish.
8. You’ll be asked if you want to change the settings
for the website, so click Yes, because you want to administer the users (see Figure 3-6).
Figure 3-8. Select Customize.
42
DEVELOPING STATIC WEBSITES
3. On the Connection Info page, select and then fill in
your SFTP settings (see Figure 3-9). The Export Assistant will now check the settings.
5. I’m going to mail Jake his key, so I leave the Send in
e-mail option selected as shown in Figure 3-10 and provide him with a password, which I will have to tell him.
Figure 3-9. Insert your SFTP details. Figure 3-10. Enter the password for the key.
4. Select Users on the Group Info page. 6. When I click the Continue button, my e-mail application opens, attaches the encrypted key, and waits for an e-mail address. As you can see in Figure 3-11, the body of the e-mail is already filled in nicely, ready to go.
Figure 3-11. Contribute words the e-mail and attaches the file.
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CHAPTER 3 Users can be allowed to alter a certain image or specific text sections, for example, all of which are definable by you. This ensures that they can’t completely destroy the layout (and all your hard work!) but don’t have to trouble you every few days when they want to update the site. There’s even a one-click command to specify that they can only edit blocks of text (see Figure 3-12). All of this works alongside Dreamweaver too. If you use templates (discussed later in this chapter), you can specify which regions you want the user to be able to edit via Contribute. The Check In/Out feature (see Figure 3-13) is in full effect too, which protects against accidental overwrites (by either party!). This feature tells you whether someone else is actually working on a file currently and, if true, prevents anyone else from making changes to that file.
Figure 3-12. Define what the user is able to alter.
Figure 3-13. The Check In/Check Out feature tells you if a file is in use.
44
DEVELOPING STATIC WEBSITES No longer will you have to trust your clients with FTP login details and worry about them deleting the whole site at 7 p.m. on a Friday night, just as you’re about to go and party for the weekend. I’m sure you’ve all had that emergency phone call, something like “I think I’ve done something wrong and the CEO needs to see the site in 5 minutes. Can you help!?” Ugh, party postponed. Not any more though!
Figure 3-14. Install the command line BBEdit tool.
BBEdit
As text editors go, BBEdit is the king. I find this application quite invaluable. It has syntax coloring for HTML, XHTML, XML, CSS, PHP4, JavaScript, Java, Perl, C/C++, Ruby, Python, Pascal, and much more, with plug-ins for other languages (such as SQL) available on its website (). It even has a glossary menu full of code snippets (from Apache through to XSLT), which you can just drag or click to insert into your document. The built-in FTP browser is also useful. Feed in your FTP settings, and you can edit documents live on the server, eliminating the need for extra FTP software. This is one speedy FTP application too! The tie-in with Dreamweaver is really cool. I often have the same page open in both Dreamweaver and BBEdit at the same time. Edit the code in one application, and it’s updated in the other one automatically. Dreamweaver hiccups a little bit and asks if you want to reload the page, but BBEdit handles itself like a champ. Opening “hidden” documents, (system files such as the Apache configuration document httpd.conf, .htaccess files, and so on) is extremely easy. In BBEdit, go to File ➤ Open Hidden and locate your file, or you can use the command line. When you first run the application, you’re asked whether you want to install the command line tool as shown in Figure 3-14. Do so, as this is a very useful feature.
Figure 3-15. Vim command line text editor
If you’re scared by the command line text editor Vim, shown in Figure 3-15 (and I don’t know anyone who isn’t, at first), and don’t really want to learn about Pico or Emacs (see Figures 3-16 and 3-17) just yet, then all you have to do is open Terminal and type in bbedit /etc/httpd/httpd.conf BBEdit will open the file. If there are Root User permissions attached to the file, you can just authorize with your Admin password, amend the file, and then reenter the password to save the document. Easy.
45
CHAPTER 3
Figure 3-16. Pico
Figure 3-17. Emacs
World Wide Web Consortium (W3C)
“The World Wide Web Consortium () was created in October 1994 to lead the World Wide Web to its full potential by developing common protocols that promote its evolution and ensure its interoperability.”
That’s what their website says, but what does that really mean? How does that affect you as a web developer? The W3C say that their long-term goals for the web are Universal access: To make the web accessible to all by promoting technologies that take into account the vast differences in culture, language,
In a nutshell, they’re trying to ensure that the web has standards and that these standards are adhered to. I mentioned earlier how browsers didn’t use to work in the same way as each other. Some would display certain types of CSS, others none. You could use certain code in one browser, but have no chance with the other browser. Most designers would get around this using the methods I described at the start of this chapter, or by creating sites entirely in Flash. While all that Flash work looked very nice, it didn’t exactly go down very well in a text-only browser. This wasn’t playing fair to people with disabilities either. How could you make the text bigger, if it was all embedded in Flash? How could you navigate the site in the usual way? All Flash sites rendered the browser’s Back/Forward buttons pretty much useless, including the keyboard shortcuts for them. This had to stop. Standards were necessary, and soon. Jeffrey Zeldman () cofounded The Web Standards Project (WaSP, at) in 1998. Working closely to the W3C’s new standards, WaSP aimed to convince Microsoft, Netscape, and other browser developers to make their new browsers comply with these standards, as opposed to each company’s individual standards. After all, what’s the point of having standards, if nobody sticks to them?
46
DEVELOPING STATIC WEBSITES The idea was that by adhering to such standards, a designer could code a page in such a way that it would render correctly in every browser on every platform. That platform could be IE 5.5 on a PC, IE 5 on a Mac, or Lynx on a UNIX system. Even a PDA or a mobile phone should produce the desired result. Standards were in place, but only a handful of designers knew about them or how to implement them. Even today, probably less than 1 percent of websites are standards compliant. Hopefully, by the time you’ve read this book, you’ll know how to make your sites comply with those standards, as I’ve tried to make sure all the code included is totally complaint to XHTML and CSS rules (which have possibly been updated somewhat, depending on when you read this book). mentioned Flash sites before, and how a visually impaired person couldn’t get the same experience out of them. Now, however, Flash MX 2004 has a whole set of accessible components to use, as well as new support materials on accessibility at the Macromedia Accessibility Resource Center for designers and developers (). If you think you’re ready to comply, and you want to test your sites, you can get Bobby ( .watchfire.com) to give them the once-over for you. Bobby will check your site against the latest WAI and/or Section 508 guidelines.
Accessibility
“The power of the Web is in its universality. Access by everyone regardless of disability is an essential aspect.” Tim Berners-Lee, W3C Director and inventor of the World Wide Web
XHTML
What is XHTML, and why are we using it? XHTML (which stands for Extensible Hypertext Markup Language) sounds cool, because it starts with an X. In real life, however, it’s just a set of document types to extend HTML, which wasn’t really designed for the kinds of things that people ended up using it for. Tags got misused, and people were applying all kinds of hacks to it to get results. It was aging fast, and other technologies were moving faster. With the ability to view web pages on a mobile phone or PDA came the ability to see how badly designed most of the Internet was. Shonky coding meant bad rendering on these new gadgets (if the page rendered at all).
Accessibility is a word you should be familiar with by now. It’s all about opening up the Internet to give everybody access to the same information. With the Internet rapidly replacing more traditional information resources such as libraries, people who are housebound because of a disability now have access to more information than ever before. Organizations like Web Accessibility Initiative (WAI, at) are committed to pursuing accessibility, to further opening up the Internet to people with disabilities. Also, the American government introduced something called Section 508 () to deal with
Enter XML
XML (short for Extensible Markup Language) was written with the Internet in mind, yet could be understood by any gadgetry that was XML-savvy. Hailed as the next Best Thing Ever™, it was designed to be completely usable and easy to read, yet powerful and easy to write. Because it used a similar format to HTML (tags, attributes, and so on), it wasn’t a completely alien language
47
CHAPTER 3 from the beginning. Open any of your preference files (those .plist files in your Preferences folder) and you’re reading an XML document. Whereas HTML was limited to a certain number of tags, XML isn’t. In fact, people often mistakenly refer to it as “Extendible Markup Language” because there are no such constraints in place. That’s not to say it doesn’t have rules. It’s far stricter than HTML, and you can’t bend the rules this time. XML is here to stay, but we’re not totally ready for it yet; we need a period of adjustment. XHTML 1.0 Strict is, as the name suggests, strict. You might like to think of it as having to take your driving test again, after all those bad habit-forming years. XHTML 1.0 Transitional is far closer to the HTML that makes up the Internet that you’re currently used to. This is probably the easiest route to take if you’re planning on converting a lot of old files, or if your current design isn’t up to Strict standards. XHTML 1.0 Frameset speaks for itself. If you’re using frames, you have to use this DOCTYPE.
This isn’t a book about XML so let’s get on with it. XHTML is a combination of XML and HTML (I bet you never saw that one coming), designed to make things as forward compatible as possible, which means sticking to standards. XHTML documents conform to the same rules as XML, as well as most of the more commonly used HTML rules. You can’t just expect modern browsers to display XML output overnight, so “transitional” is the best way of describing the period we’re in, hence combining the two to produce XHTML. Even though we’re only up to version 1.1 of XHTML, it’ll be written in such a way that all future versions should be able to read 1.1 forever. As the latest versions of GoLive and Dreamweaver are XHTML compliant, you don’t have to worry about your code.
Namespace
Next up, you need an xmlns (XML namespace) declaration. XML namespaces allow qualifying elements and attribute names used in the document by associating them with namespaces identified at the location shown: <html xmlns=" ➥xhtml" xml:
In XHTML, you must write all tags in lowercase. XML is case sensitive; therefore <div> is completely different than <DIV>, and so on. While in HTML certain tags didn’t have a closing tag, XML doesn’t allow such reckless behavior. As such, you have to close every tag. For example: <li>this list item needs closing</li> <p>that goes for paragraph tags too</p> Even the empty tags need to be closed. <br> is now <br /> (note the space before the slash). <img src="artley.jpg" width="320" ➥ Be sure to wrap those values in quotes too! Incorrect: width=320. Correct: width="320".
DOCTYPE
Because you’re now conforming to XML rules these days, there are certain things you must have in place in order for a document to be classed as valid XHTML. The first thing you need is the DOCTYPE declaration. This is one of three varieties: <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML ➥1.0 Strict//EN" " ➥xhtml1/DTD/xhtml1-strict.dtd"> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML ➥1.0 Transitional//EN" " ➥TR/xhtml1/DTD/xhtml1-strict.dtd"> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML ➥1.0 Frameset//EN" " ➥xhtml1/DTD/xhtml1-strict.dtd">
48
DEVELOPING STATIC WEBSITES
Div
A div is short for “division.” For example, having content grouped together is a division of your HTML document. You’ll be meeting the <div> tag at the end of this chapter, when you come to lay out some code.
import
The @import method only works in browser versions 5.0 and upwards, so using @import would effectively “hide” the sheet from 4.0 browsers. <style type="text/css" media="all"> ➥@import "fancy_style.css";</style>
Cascading Style Sheets
They’ve been around for a while, but Cascading Style Sheets aren’t just being used to change a few font colors any more. There is far more to them than that, which I’ll briefly touch on here. Previously, page layout has been driven by the <table> tag. If you view source code on an average site, you’re likely to see a massive expanse of <tr> and <td> tags littering the code. Not only did this create some seriously long and often messy code, but technically it was invalid markup, most of the time. Oh, did I mention that you’d have sleepless nights, trying to get the same result in each browser? And then there are the cheap hacks like a transparent shim.gif to get a table cell to be a certain width or height. If you have the desire to become a CSS Grandmaster, I suggest grabbing a book entirely about CSS such as Cascading Style Sheets: Separating Content from Presentation, Second Edition by Owen Briggs et al. (Apress, 2004). You can also check out what the W3C have to say at.
embed
You might only have a few CSS rules, or you might have a couple that you only need to apply to one page. If this is the case, you can embed them in the head of your XHTML document like this: <style type="text/css"> <!-h1 { font-family: verdana, arial, Helvetica, ➥sans-serif; color: #f00; font-weight: bold; } --> </style>
CSS basics
Let’s have a basic look at some CSS styles. With each rule, we’re dealing with two parts: a selector and a declaration. In action, that looks like this: h1 {color: #f00;}
Linkage
Before your XHTML document can read from the style sheet you’re going to write, it has to know where to find the style sheet. You can do this in a number of ways.
h1 is the selector and the contents of the curly braces are the declaration. You’ve told your document to make all h1 tags red. I could have said color: #red or color: #ff0000, but shorthand saves on typing, and that’s always a good thing to get into the swing of.
link
The standard way to link a style sheet is . . . the link method! <link rel="stylesheet" type="text/css" ➥ You can use “shorthand” to save on both typing and file size. As each pair of numbers corresponds to R, G, and B values, you can combine them if a pair is the same. So, for #006633, you can simply use #063 instead. You can’t, however, use shorthand for a value such as #5c892e.
49
CHAPTER 3 The <font> tag has been deprecated. Gone are the days of bloated code, rammed full of lines like <h1><font face="verdana, arial, ➥helvetica, san-serif" size="2" color= ➥"#FF000"> witty line of text!</font></h1> Today, this is simply replaced by a line in your style sheet, like so: h1 { font-family: verdana, arial, Helvetica, ➥sans-serif; color: #f00; } To add further formatting to this, you could make it bold as follows: h1 { font-family: verdana, arial, Helvetica, ➥sans-serif; color: #f00; font-weight: bold; } If you wanted to apply this to all six of the header tags, you could do that simply enough. h1, h2, h3, h4, h5, h6 { font-family: verdana, arial, Helvetica, ➥sans-serif; color: #f00; font-weight: bold; } You could then determine the size of each one in individual statements as follows: h1 { font-size: 32px; } h2 { font-size: 24px; } . . . and so on. The <body> tag in an XHTML document is usually just that: <body>. All your <body> elements are now contained within CSS. Here’s an example: body { background: #ccc url(background.gif) ➥top right no-repeat fixed; color: #000; margin: 0; padding: 0; border: 0; } What you did there was set the body background color to #ccc (a light gray), with any text set to #000 (black). Instead of using leftmargin="0" topmargin="0" marginwidth="0" marginheight="0" to set those margins to zero, you now use margin: 0; (with the same applying to padding). Back to that first line—notice there’s an image there. There’s no need to make GIFs about 5000 pixels high now in order for them to seem like they don’t scroll (that was always a cheap—but necessary!—hack anyway). Now you can just specify which position you want the image to be in, whether it should be fixed, and whether it should tile. You can even specify x-axis tiling only and likewise for y-axis tiling. Much better than a cheesy old hack job. Comments in CSS look like this: /* this is a one-line CSS comment */ /* you can also split comments over a few lines like this */
CSSEdit
Of course, you don’t have do all of this by guesswork and straightforward typing. You can use Dreamweaver, GoLive, or a shareware application called CSSEdit ($24.99, from MacRabbit at cssedit), shown in Figure 3-18. Installing it is child’s play, as with a lot of Mac apps. Simply mount the disc image and then drag the icon into the Applications folder. Done.
50
DEVELOPING STATIC WEBSITES
My first style sheet
Now that you’ve had a quick look at how and why things work, it’s time to get stuck into some code. You’re not aiming to win any design awards here, just write some valid markup code, learn about the subject matter, and output something that you can feel a little bit proud of. Something like the page in Figure 3-20. Get your text editor ready for action, open a new Finder window, and let’s rattle some ASCII.
1. First, make yourself the following directories,
because the files you’ll be generating are going to need some form of structure:
Figure 3-18. CSSEdit is worth its weight in gold.
~/username/Sites/chapter3 ~/username/Sites/chapter3/css ~/username/Sites/chapter3/images ~/username/Sites/chapter3/js
Point, click, and see a live preview as shown in Figure 3-19. It’s a very nice little application, and certainly worth the shareware fee.
2. Create a new text file called green.css and save it
in the /css directory. Okay, you need some body elements first, so type in the following code: /* green scheme */ body { background: #ccc; color: #000; margin: 0; padding: 0; border: 0; } #ccc gives you a gray background, which will be outside of your container, with #000 giving you a black text color should any text escape onto the body.
3. Next, let’s tackle the fonts. Carrying on directly
underneath the previous code, type the following: /* main page fonts */ p { margin-top: 0; margin-bottom: 1em; font: 11px/16px "Lucida Grande", ➥"Lucida Sans Unicode", verdana, lucida, ➥arial, helvetica, sans-serif; } code continues
Figure 3-19. As you can see, CSSEdit makes things a lot easier to visualize, as opposed to just “code and hope.”
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CHAPTER 3
Figure 3-20. This is how the page you are about to code should look, at the end of this chapter.
a:link, a:visited { font-weight : bold; text-decoration : none; color: #84a563; background: transparent; } a:hover { font-weight : bold; text-decoration : underline; color: #557239; background: transparent; } a:active { font-weight : bold; text-decoration : none; color: #5c892e; background: transparent; }
Here, you’ve specified no underline for links unless the cursor is hovering over it.
4. Now on to the header tags:
/* header fonts */ h1, h2, h3, h4, h5, h6 { font-weight: normal; font-family: "American Typewriter", ➥"Trebuchet MS", Trebuchet, Lucida, ➥sans-serif; } h1 { margin-top: 0; margin-bottom: 0; font-size: 32px; text-transform: lowercase; }
52
DEVELOPING STATIC WEBSITES h2 { font-size: 24px; margin-top: 25px; margin-bottom: 0; letter-spacing: 1px; } { font-size: 16px; margin-top: 0; margin-bottom: 0; } { font-size: 13px; margin: 5px 0; padding: 0; letter-spacing: 1px; }
1. Continuing at the bottom of the same green.css
document, let’s go wild and make a box on the screen (close the curtains in case the neighbors can see this frenzy of activity, they’ll only get jealous). /* main layout divs */ #container { background: #fff; border: 1px dotted #000; margin: 10px; padding: 10px; width: 700px; } Let’s dissect that, shall we? #container is the name you’ve given to this element, because it will eventually contain the rest of your content. background: #fff; sets the background of the box to white. border: 1px dotted #000; gives your box a black dotted border that’s 1 pixel wide. margin: 10px; sets the space outside of your box. padding: 10px; sets the space inside of your box. width: 700px; sets the width of the box. (Cunning, eh?) margin and padding actually have four values there, but you can utilize a shorthand rule to knock them down to one value of 10 pixels. The longhand version would be 10px 10px 10px 10px; The order can be easily remembered, if you think in a clockwise direction starting at 12 o’clock: top, right, bottom, left. So, if you just wanted a margin of 10 pixels on the right, that would be 0px 0px 0px 10px;. Once again, there is a lot more shorthand, but this isn’t a 1000-page CSS manual.
h3
h4
As all of the header tags are using the same font, you can group them all together. After doing that, you specify the individual sizes. So far, there probably hasn’t been too much that you haven’t seen before. Fonts, padding, sizes: it’s probably all familiar stuff. That feeling of familiarity is about to leave, so wave goodbye.
Positioning
One of the most abused tags in HTML was probably the <table> tag. Created originally to contain tabular data, this poor tag was dragged all over the place. Yes, you could get your layout looking like you wanted it but not only was it invalid markup, it was also likely to look awful in whichever browser you forgot to check it in (that just so happens to be the client’s favorite browser . . . doh!). This doesn’t mean that the <table> tag has gone the same route as the <font> tag, by any means. You can still use it, when used correctly (i.e., for tabular data). To lay out your document without tables may seem scary at first, if this is what you’re used to, but it’s nowhere nearly as scary once it’s been explained to you. It’s just a matter of using a few different lines of CSS.
If you want the container centered, replace margin: 10px with this line: margin:10px auto;.
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CHAPTER 3 Not so tough to get to grips with, hmm? Hang on; You just need to clear up some Internet Explorer mess. Thanks to Tantek Çelik’s Box Model Hack (www .tantek.com), you can do that with a few lines of code. There you have it. You’ve got a box, ready for content as soon as you start your XHTML document. Let’s kick out the rest of the CSS quickly, so we can get on with coding that page.
2. If you head back to that width value (width:
700px), you need to hack it up a bit. Some earlier versions of PC Internet Explorer will make a mess of the padding value. You need to fix that, so use the following new code instead (highlighted in bold): #container { background: #fff; border: 1px dotted #000; margin: 10px; padding: 10px; width: 722px; /* false value ➥for IE4-5.x/Win, like so: */ voice-family: "\"}\"""; /* real width ➥+ l/r border + l/r padding = false value */ voice-family:inherit; /* end false ➥value for IE4-5.x/Win */ width: 700px; /* real value ➥for compliant browsers */ } What this does is trick IE into thinking that you want your box to be 722 pixels wide (which you do, ultimately). This value is determined by adding together the following: The real value + any left/right borders + left/right padding. PC IE 5.x doesn’t understand voice-family rules, so anything after that will go unnoticed. Once it’s been tricked into giving you a workable value, the next two lines get it drunk enough not to notice the real value, which you then state while it’s not looking.
4. Let’s add a banner at the top of the page. You’ll lay
this out in a similar style to the container you just made. #banner { padding: 2px 2px 2px 10px; background-color: #84a563; font-family: "Lucida Grande", "Lucida ➥Sans Unicode", verdana, lucida, ➥sans-serif; font-size: 18px; color: #fff; font-weight: bold; letter-spacing: 0.33em; text-align: left; }
5. You’re moving down the page, from the left, so it’s
time to bang that sidebar into position. As you’re specifying a width, you need to include those IE/Opera workarounds too. #sidebar { float: left; margin: 0; margin-right: 2px; padding: 2px; background-color: #84a563; font-family: "Lucida Grande", "Lucida ➥Sans Unicode", verdana, lucida, ➥sans-serif; font-size: 11px; font-weight: normal; color: #fff; text-decoration: none; width: 149px; /* false value ➥for IE4-5.x/Win, like so: */ voice-family: "\"}\""; /* real width + ➥l/r border + l/r padding = false value */ voice-family:inherit; /* end false ➥value for IE4-5.x/Win */ width: 145px; /* real value ➥for compliant browsers */ }
3. One final hack needs adding, which is commonly
referred to as the “Be nice to Opera” rule. This is just html>#container { width: 700px; ➥to Opera */ } /* be nice
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DEVELOPING STATIC WEBSITES html>#sidebar { width: 145px; } All you’re doing here is simply altering the properties of an unordered list (<ul>). You should be able to tell what’s going on by now, apart from one line. Normal lists have bullet points, numbers, and so on, but you don’t want any of that stuff here, so you use list-style-type: none; to get rid of that clutter.
/* be nice to Opera */
6. Moving right from the sidebar, you want somewhere for the actual content to live, so create a <div> for it like this: #content { padding: 2px; margin-left: 158px; background-color: #fff; font-family: "Lucida Grande", "Lucida ➥Sans Unicode", verdana, lucida, ➥sans-serif; color: #000; font-size: 12px; } As you can see, you’ve told #content <div> to start 158 pixels in from the left, but feel free to experiment with this distance. Use a smaller figure and you’ll see some funk occurring. The more you play around with these things, the more you’ll understand how they work. Don’t just take my word for it; push things until they break, then fix them. Now you have almost all of your main areas set out. As well as working from left to right, you’ve obviously been working from top to bottom too. You do this so that things are easier to read when skimming through the green.css document. You could have the #header <div> detailed at the bottom of the CSS document, but it makes far more sense to be logical and place those elements appearing at the top of the page at the top of your code. With this in mind, you won’t be dealing with the page footer for a while.
8. Now add the list item properties.
#navcontainer li { margin: 0 0 3px 0; } This gives you a 3-pixel gap underneath each list item.
If you want a rollover image, uncomment the background-image line in the following #navcontainer code in step 9 and make an appropriate image.
9. Okay, you’ve specified that your list is going to be
unordered, and provided the list items with a nice gap underneath. Next, you’ll create the actual containers that the user will click on. Again, you’re specifying a width, so it’s time to wheel in the IE/Opera show. #navcontainer a { display: block; padding: 2px 2px 2px 20px; border: 1px dotted #fff; width: 100px; height: 20px; background-color: #5c892e; /* background-image: ➥url(images/roll_down.jpg); */ width: 124px; /* false value ➥for IE4-5.x/Win, like so: */ voice-family: "\"}\""; /* real width + ➥l/r border + l/r padding = false value */ voice-family:inherit; /* end false ➥value for IE4-5.x/Win */ width: 100px; /* real value ➥for compliant browsers */ } html>#navcontainer { width: 100px; ➥to Opera */ }
7. The links in the sidebar are next.
/* sidebar links */ #navcontainer ul { margin: 0; padding: 0 0 0 10px; list-style-type: none; font: normal 10px/18px "Lucida Grande", ➥"Lucida Sans Unicode", verdana, lucida, ➥sans-serif; }
/* be nice
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10. Now take care of the fonts:
#navcontainer a:link, #navlist a:visited { font-size: 10px; color: #eee; text-decoration: none; }
#styletool img { display: inline; border: 0; padding: 2px; }
13. Underneath the style selector buttons, you’re going
to show the majority of the Internet that you are, in fact, compliant with these new web standards. Again, these are to be images, but they’re still in need of some layout guidance. /* validation buttons */ #valid { border: 0; margin: 10px 0 15px 20px; } #valid img { display: inline; border: 0; padding: 2px 2px 2px 10px; }
11. Let’s give the user that “By the power vested in me,
I can make things change color!” feeling, and tell things to get excited when the mouse is on the prowl. #navcontainer a:hover { font-size: 10px; border: 1px dotted #fff; background-color: #557239; /* background-image: ➥url(images/roll_over.jpg); */ color: #fff; } #active a:link, #active a:visited, ➥#active a:hover { font-size: 10px; border: 1px dotted #fff; background-color: #557239; /* background-image: ➥url(images/roll_over.jpg); */ color: #fff; }
14. Lastly, add the footer, which plays by the same
IE/Opera rules as the other main containers. /* footer */ #footer { clear: both; padding: 2px; margin-top: 2px; background-color: #84a563; font-family: "Lucida Grande", ➥"Lucida Sans Unicode", verdana, ➥lucida, sans-serif; font-size: 9px; font-weight: normal; text-transform: lowercase; color: #fff; text-align: right; }
12. As this is to be an all-compliant accessibility exercise, you’re going to give the user the opportunity to change some aspects of the page. This will be a change of colors, as well as font sizes, both of which benefit the visually impaired. You’ll use some images to show what these buttons will do, but they still need positioning. /* the style selector */ #styletool { border: 0; margin: 10px 0 15px 0; padding: 0 0 0 9px; }
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DEVELOPING STATIC WEBSITES This file should be saved as /css/green.css. The easiest way to make your alternative color sheet is to then save this file as /css/purple.css, and then use BBEdit to run a find-and-replace operation on the actual colors. You can do the same for the large type and reversed color sheets too, making sure to add a few pixels to those font sizes. Save these sheets as /css/large.css and /css/reverse.css. The following sections detail the schemes that I came up with (included in the download files), but feel free to choose your own colors.
Green color scheme
This file should be saved as /css/green.css (see Figure 3-21). #84a563; is used for a:link, a:visited, #banner, #sidebar, and #footer. #5c892e; is used for a:active and #navcontainer a. #557239; is used for a:hover, #navcontainer a:hover, #active a:link, #active a:visited, and #active a:hover.
Figure 3-21. If all went according to plan, you should be looking at a page like this now.
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Purple color scheme
To create this alternative color scheme style sheet, follow these steps:
3. The easiest way to do this is with the Find and
Replace function in BBEdit. Go to Search ➤ Find to open the Find and Replace window.
1. First save /css/green.css as /css/purple.css. 2. You need to substitute the colors that you used in
the green.css file for the following color scheme: #838; for a:link, a:visited, a:hover, #banner, #sidebar, and #footer. #606; for a:active and #navcontainer a. #c6c; for #navcontainer a:hover, #active a:link, #active a:visited, and #active a:hover.
4. In the top Search For field, type in the existing
color you want to change, and type the new color you want to change it to in the bottom Replace With field. For example, to change a:active and #navcontainer a from the color #5c892e to #606, fill in the window as shown in Figure 3-22.
Figure 3-22. Using Find and Replace in BBEdit. Note the Start at Top option!
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5. Click the Replace All button, and your two
BBEdit will search the document from wherever the cursor is currently, to the end of the document. To avoid this, place the cursor at the start of the document. You can also select the Start at Top option in the Search window. To turn this on permanently, open BBEdit’s preferences, and then select the Text Search pane. Check the Remember Find dialog box’s Start at Top settings. instances of the color should be updated in your new purple style sheet (see Figure 3-23).
Figure 3-23. The dialog box tells you how many occurrences it replaced.
Figure 3-24 shows what the page will look like with the purple color scheme.
Figure 3-24. Click the purple switcher button, and you get this view.
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Large type scheme
Now move on and create the large type scheme style sheet (/css/large.css). Aside from the colors, the only change here is the font size. I’ve just added 4 pixels to each font size in the document. Here’s the new color scheme (see Figure 3-25): #999; for a:link, a:visited, #banner, #sidebar, and #footer. #666; for a:active and #navcontainer a. #333; for a:hover, #navcontainer a:hover, #active a:link, #active a:visited, and #active a:hover.
Reversed color sheet
As you can see, this scheme (/css/large.css) is a bit different. It’s for partially sighted people, most of whom feel more comfortable with light text on dark backgrounds. The font size has also been bumped up a few more notches yet again, adding 10 pixels to the original size. The links and text colors are all #ff9, and the grays are the same as the large type scheme. See if you can match them up (refer to Figure 3-26).
Figure 3-25. Click the large type switcher button, and you get this view.
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Figure 3-26. Click the reverse colors switcher button, and you get this view.
JavaScript
As with a few things I’ve mentioned in this chapter, this isn’t the time to go into huge details. One thing falling squarely into that category is JavaScript. If you’ve ever created a rollover, then you’ve probably used JavaScript. You might have generated that JavaScript via Dreamweaver or Fireworks, but it was there all the same. Bandwidth usage, and how to save on it, has already been touched on, but there’s another way you can cut down on the code bloat. The average Macromedia rollover effect produces something like this in the head of your document: <script language="JavaScript" ➥ <!-function MM_preloadImages() { //v3.0 var d=document; if(d.images){ ➥if(!d.MM_p) d.MM_p=new Array(); var i,j=d.MM_p.length, ➥a=MM_preloadImages.arguments; ➥for(i=0; i<a.length; i++) if (a[i].indexOf("#")!=0){ ➥d.MM_p[j]=new Image; ➥d.MM_p[j++].src=a[i];}} } (code continues)
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CHAPTER 3 function MM_swapImgRestore() { //v3.0 var i,x,a=document.MM_sr; ➥for(i=0;a&&i<a.length&&(x=a[i]) ➥&&x.oSrc;i++) x.src=x.oSrc; } function MM_findObj(n, d) { //v4.01 var p,i,x; if(!d) d=document; ➥if((p=n.indexOf("?")) ➥>0&&parent.frames.length) { d=parent.frames[n.substring(p+1)] ➥.document; n=n.substring(0,p);} if(!(x=d[n])&&d.all) x=d.all[n]; for ➥(i=0;!x&&i<d.forms.length;i++) ➥x=d.forms[i][n]; for(i=0;!x&&d.layers&&i ➥<d.layers.length;i++) x=MM_findObj ➥(n,d.layers[i].document); if(!x && d.getElementById) ➥x=d.getElementById(n); return x; } function MM_swapImage() { //v3.0 var i,j=0,x,a=MM_swapImage.arguments; ➥document.MM_sr=new Array; ➥for(i=0;i<(a.length-2);i+=3) if ((x=MM_findObj(a[i]))!=null) ➥{document.MM_sr[j++]=x; ➥if(!x.oSrc) x.oSrc=x.src; x.src=a[i+2];} } //--> </script> If you have the same rollovers on 50 or more pages of a site, then this is just a ridiculous waste of time and bandwidth. Fortunately, there’s an easy way to tackle this. This way, you’re cutting out all the boated code from each page, and the user only has to download the file once. Everybody wins! While you’ll be using a JavaScript style switcher in this example to change the style sheets, I didn’t write it myself (gasp) because in the wonderful world of web development, people are nice enough to write things and willingly let you borrow them. That is how I came by the styleswitcher.js file you’ll use in this example. You can download the code with an explanation of what does what from the following sites: Save this file to your chapter3/js folder. If you want to investigate JavaScript in more depth, there are many resources on the web. is a good place to start.
Status bar message
In the bottom left of most browser windows, you can usually see some text whizzing past, as the browser loads a page (Safari breaks with this tradition). This is called the status bar, and you can use JavaScript to have a default status bar message there. Just expand your <body> tag to read like this: <body onload="window.defaultStatus=' ➥friends of ED | Mac OS X Web Development ➥Fundamentals: Chapter 3'">
1. Cut all of this code and paste it into a new document.
Templates and Library Items
When you have a heavy workload, saving time whenever possible is essential. Dreamweaver’s templates and Library Items can seriously cut down time in a number of ways. For starters, once you’ve decided on your page layout, you can just turn that page into a template and keep reusing it for every new page in your site. But how do these work? Well, you set “editable regions” where your content is going to be changing for each page. Typically, this could be the page title, possibly a
2. Save the new document as global.js in your /js
directory.
3. Link it in the <head> of all those pages, like this:
<script language="JavaScript" ➥</script>
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DEVELOPING STATIC WEBSITES banner graphic, and, of course, the page content. This is useful if you’re not working alone on a project, as you can rest assured that the rest of the team are all working from the same template. If you use Contribute (as discussed earlier in this chapter), you can rest easy there too. When your client logs onto the server to alter any pages with Contribute, the application honors those same editable regions. There’s no way clients can delete anything they’re not supposed to. Library Items are useful if you’re designing a site with certain assets on every page, which may be subject to change before the site goes live. Rather than wait for the final assets to turn up, you could use a placeholder image, turn it into a Library Item, and then drag this onto your documents. This way, all you have to do when you get the final assets is update the Library Item, and this will update through all the pages you linked it from. Just make sure to upload the /Library folder to the server or you’ll be in big trouble. As you can see, you’re using XHTML 1.0 Transitional rather than Strict 1.0, because it’s more than suitable for your needs. Once you’ve got your head around this, you can venture forth into the world of Strict 1.0 forever. You’ve also added the CSS document you made earlier in this chapter.
2. Now you need some structure for your content, so
position your <div> tags. To avoid confusion, these will be added a couple at a time. These go between the <body> tags, like so: <body onload="window.defaultStatus=' ➥friends of ED | Mac OS X Web Development ➥Fundamentals: Chapter 3'"> <div id="container"> <div id="banner">this is the #banner ➥<div><div> </div> <!-- end of the container div --> </body> As you might have guessed, this sets out the main area, which will contain all the content, and the banner <div> up at the top. I’ll comment the end tags, so you have a clear idea of which tag is doing what. Each <div id> corresponds to a section of the style sheet. So, <div id="container"> gets its instructions from the #container section of the style sheet. You can see how this works by just unlinking the style sheet and watching how the page lays itself out (similar to Netscape 4.x!).
XHTML time
Okay, you’ve got your Cascading Style Sheet all nicely typed out, so let’s go ahead and finish coding that page.
1. Start a new text file and save it as index.html in
the root directory of your site folder structure. Begin by typing in this basic starting block of code: <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML ➥1.0 Transitional//EN" " ➥TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns=" ➥xhtml" xml: <head> <meta http- <title>Chapter 3 Example</title> <link rel="stylesheet" type="text/css" ➥</head> <body> </body> </html>
3. Next, add these <div>s between the previous pair:
<div id="banner">this is the #banner ➥<div><div> <div id="sidebar"> <div id="navcontainer"> </div> <!-- end of the navcontainer div --> </div> <!-- end of the sidebar div --> </div> <!-- end of the container div --> Unsurprisingly, this sets up the sidebar on the left of the page, where your navigation container will sit. Let’s add some links to it.
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4. Add the following code inside the navcontainer
<div>: <div id="navcontainer"> <ul id="navlist"> <li id=" active "><a href="index.html" ➥Home</a></li> <li><a href="link2.html" title="Link 2 ➥title">Link 2</a></li> <li><a href="link3.html" title="Link 3 ➥title">Link 3</a></li> <li><a href="link4.html" title="Link 4 ➥title">Link 4</a></li> <li><a href="link5.html" title="Link 5 ➥title">Link 5</a></li></ul> </div> <!-- end of the navcontainer div --> The first link here has an ID called active. This simply states which page you’re on now, so the color for that link is different. On your Link 2 page, you’d move that ID down to Link 2, and so on.
➥ <!-- the @import method only works from ➥5.0 and upwards --> <!-- so, using @import would "hide" the ➥more sophisticated sheet ➥from < 5.0 browsers --> <!-- <style type="text/css" media="all"> ➥@import "fancy_style.css";</style> --> <script language="JavaScript" ➥</script> </head> The commented-out @import section shows you how you would include a .css file that wouldn’t work in browsers earlier than version 5.0, so, using this method, those earlier browsers wouldn’t even read this line. Maybe you have some cutting edge CSS3 code that you want to show off? Only the very latest browser will be able to handle it, so import it using this method. Back to the <body> of the document, and it’s time to code those links. The way you’re calling them is as follows: <a href="#" title="Switch Styles: green" ➥ ➥<img src="images/selector_green.gif" ➥</a> It’s a basic href link with both a title and an access key (which will be discussed in the next chapter) in keeping with those standards. The call to the JavaScript is setActiveStyleSheet, which then tells the browser which style sheet to select.
5. Next, let’s add those style buttons. As with the
global.js document mentioned earlier, you first need to attach the JavaScript file to your document. You do this by adding the following line within the <head> tags like this: <title>Chapter 3 Example</title> <link rel="stylesheet" type="text/css" ➥ <script language="JavaScript" ➥</script> </head>
6. Because you’re using alternative style sheets on
this job, you need to have those sheets specified as such. Here’s how you do it: <title>Chapter 3 Example</title> <link rel="stylesheet" type="text/css" ➥ <link rel="alternate stylesheet" ➥ <link rel="alternate stylesheet" ➥ <link rel="alternate stylesheet"
7. Add this to the bottom of your existing code:
<li><a href="link5.html" ➥Link 5</a></li> </ul> </div> <!-- end of the navcontainer div --> <div id="styletool"> <a href="#" title="Switch Styles: ➥green" onclick="setActiveStyleSheet ➥('green'); return false;" accesskey="g"> ➥<img src="images/selector_green.gif" ➥</a>
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DEVELOPING STATIC WEBSITES <a href="#" title="Switch Styles: ➥purple" onclick="setActiveStyleSheet ➥('purple'); return false;" accesskey="p"> ➥<img src="images/selector_purple.gif" ➥</a> <a href="#" title="Switch Styles: ➥large type" onclick="setActiveStyleSheet ➥('large'); return false;" accesskey="l"> ➥<img src="images/selector_large.gif" ➥</a> <a href="#" title="Switch Styles: ➥reverse colors, large type" onclick=" ➥setActiveStyleSheet('reverse'); ➥return false;" accesskey="r"><img src=" ➥images/selector_large_reverse.gif" ➥</a> </div> </div> <!-- end of the sidebar div --> </div> <!-- end of the container div --> The buttons used in this example are from Taylor McKnight’s Steal These Buttons site (http:// gtmcknight.com/buttons). They are also in the download files for this chapter. To use them in this exercise, copy them to your chapter3/images folder. </div> <div id="valid"> <a href=" ➥referer" title="Validated XHTML 1.0"> ➥<img src="images/xhtml10.png" alt="XHTML ➥1.0 valid" width="80" height="15" ➥</a><br /> <a href=" ➥css-validator/check/referer" ➥ ➥<img src="images/css.gif" alt="CSS valid" ➥ ➥</a><br /> <a href=" ➥mynewtester/cynthia.exe?Url1= ➥" ➥ ➥<img src="images/sec508a.gif" alt="508 ➥valid" width="80" height="15" ➥</a> </div> </div> <!-- end of the sidebar div --> </div> <!-- end of the container div --> You can get these buttons from Taylor McKnight’s site () or this book’s download files. Remember to put them in your chapter3/images folder.
8. Immediately beneath the last block of code you
typed, you’re going to have a set of buttons to show people that all your code validates and meets the new standards.
9. If you save and preview this file in your browser
now, you’ll see things are beginning to shape up (see Figure 3-27)!
Figure 3-27. A preliminary check, and you should see things starting to take shape.
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CHAPTER 3 Well, almost. The container <div> has no height value and no content, other than the banner <div>, so that’s all you’ll see of that until it gets some content. So without further ado, let’s add some content in there, right after the end of the sidebar <div>. Let’s use good old Lorem Ipsum for now. You can grab this filler text from, and you might want to repeat it a few times, to expand the content <div>, so you can see how things expand..</p> <p>Ut wisi enim ad minim veniam, quis ➥nostrud exerci tation ullamcorper ➥suscipit lobortis nisl ut aliquip ex ea ➥commodo consequat. Duis autem vel eum ➥iriure dolor in hendrerit in vulputate ➥velit esse molestie consequat, vel illum ➥dolore eu feugiat nulla facilisis at ➥vero eros et accumsan et iusto odio ➥dignissim qui blandit praesent luptatum ➥zzril delenit augue duis dolore te ➥feugait nulla facilisi.</p></div> ➥<!-- end of the content div -->
10. So, immediately after the last block of code, add
the following: </div> </div> <!-- end of the sidebar div --> <div id="content"> <p>this is the #content <div></p> <p><a href="" ➥Lorem. Ut wisi ➥enim ad minim veniam, quis nostrud ➥exerci tation ullamcorper suscipit ➥lobortis nisl ut aliquip ex ea commodo ➥consequat.</p> <p>Duis autem vel eum iriure dolor in ➥hendrerit in vulputate velit esse ➥molestie consequat, vel illum dolore eu ➥feugiat nulla facilisis at vero eros et ➥accumsan et iusto odio dignissim qui ➥blandit praesent luptatum zzril delenit ➥augue duis dolore te feugait nulla ➥facilisi. ➥<a href="" ➥Lorem
11. Last but not least, complete the site by adding the
footer at the bottom of your code. </div> <!-- end of the content div --> <div id="footer">this is the #footer ➥<div><br />everything ©2003 ➥freakindesign</div> </div> <!-- end of the container div --> </body> </html>
12. When you add all this together and preview in your
browser, you should see something like the layout in Figure 3-28. If all went according to plan, you should be able to see similar results in most compliant browsers, although it still won’t look like a complete mess either, if you view it in noncompliant browsers.
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DEVELOPING STATIC WEBSITES
Figure 3-28. The finished page! Success!
Chapter review
Now that you’re at the end of the chapter, you should feel like you’re actually getting somewhere. You’ve learned something about web standards and why we have to have them. You should know what all those acronyms actually stand for now! More importantly, you’ve actually got some code under your belt, and because you wrote it by hand, you know how each part of it works. Later on, you’ll be using Dreamweaver, and you should be able to understand the code. None of the code from this example is written in stone, of course. Experiment with colors, fonts, etc., but change other things too. If you had to go over a section a few times before it sank in, then experiment with that section, to make doubly sure you understand how and why it works. “Think for yourself. Question authority.” Find the time to check your page in as many different browsers as you can, on different platforms. That should help you appreciate how appearances can change. Speaking of browsers, let’s check into Chapter 4, for a closer look.
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Joel Reymont wrote: >. I use a set of strings for the symbol table (I don't record the types of the identifiers, but you can add it back). I don't allow for whitespace, but you can add it back. The parser returns a string rather than a constructor with a string, but you can add it back. It is necessary to fuse the lexer and the parser together, so that they share state; but we can fuse them in a way that still leaves recognizable boundary, e.g., in the below, string "blah", ident, num, name, and numeric_simple are lexers (thus when you add back whitespace you know who are the suspects), and p0 is a parser that calls the lexers and do extra. The name lexer returns a sum type, so you can use its two cases to signify whether a name is in the table or not; then ident and num can fail on the wrong cases. (Alternatively, you can eliminate the sum type by copying the name code into the ident code and the num code.) import Text.ParserCombinators.Parsec import Monad(mzero) import Data.Set as Set main = do { input <- getLine ; print (runParser p0 Set.empty "stdin" input) } p0 = do { string "Output" ; string ":" ; i <- ident ; string "(" ; numeric_simple ; string ")" ; updateState (Set.insert i) ; return i } numeric_simple = many digit ident = do { n <- name ; case n of { ID i -> return i ; _ -> mzero } } name = do { c0 <- letter ; cs <- many alphaNum ; let n = c0 : cs ; table <- getState ; return (if n `Set.member` table then NUM n else ID n) } data Name = NUM String | ID String num = do { n <- name ; case n of { NUM i -> return i ; _ -> mzero } } | http://www.haskell.org/pipermail/haskell-cafe/2007-April/024068.html | CC-MAIN-2014-35 | refinedweb | 279 | 75.84 |
I have a class
class ActivationResult(object):
def __init__(self, successful : bool):
self._successful = successful
def getSuccessful(self) -> bool:
return self._successful
And a test
def testSuccessfulFromCreate(self):
target = ActivationResult(True)
self.assertEquals(target._successful, True)
self.assertEquals(target.getSuccessful, True)
The first assert is good, but the second one fails with
AssertionError: <bound method ActivationResult.getSuccess[84 chars]EB8>> != True
The same thing happens, when I try to print it. Why?
You are getting the method, not calling it.
Try :
self.assertEquals(target.getSuccessful(), True) # With parenthesss
It's OK the first time because you get the attribute
_successful, which was correctly initialized with
True.
But when you call
target.getSuccessful it gives you the method object itself, where it seems like you want to actuall call that method.
Explanation
Here is an example of the same thing that happens when you print an object's method:
class Something(object): def somefunction(arg1, arg2=False): print "Hello SO!" return 42
We have a class, with a method.
Now if we print it, but not calling it :
s = Something() print s.somefunction # NO parentheses >>> <bound method Something.somefunction of <__main__.Something object at 0x7fd27bb19110>>
We get the same output
<bound method ... at 0x...> as in your issue. This is just how the method is represented when printed itself.
Now if we print it and actually call it:
s = Something() print s.somefunction() # WITH parentheses >>>Hello SO! >>>42
The method is called (it prints
Hello SO!), and its return is printed too (
42) | http://m.dlxedu.com/m/askdetail/3/ab9f60076fd7b150fc6ce8c107af7a66.html | CC-MAIN-2018-22 | refinedweb | 252 | 52.76 |
...
Course Schedule Course Modules Review and Practice Exam Preparation Resources
Assignment 1
Note: Prepare the answers to these assignment questions in Word and save them as one Word document on
your hard drive. See the TX2 Assignment Submission/FAQ section for the recommended format and filename.
When your file is complete and you are ready to submit it for marking, select your “TX2 Assignment
Submission” section under the “My CGA” tab.
Question 1 (20 marks)
Important: Multiple-choice questions are to be completed within the Online Learning Environment in your
TX2 Assignment Submission section. This portion of the assignment will be automatically graded. Do not
include your answers in your Word document as they will not be graded.
Multiple choice (2 marks each)
a. Marc is the son-in-law of Bernard. Bernard owns all the issued shares of Mammoth Inc., an
import company whose taxation year ends December 31. In September 2011, Mammoth Inc.
granted Marc a $10,000 loan, bearing interest at the market rate in effect at that time, so that
he could purchase a snowmobile. Assuming that the loan is repaid in 2013 and Marc has always
made his interest payments on time, what will be the tax consequences of this loan?
1. None.
2. Under section 80.4, Marc must include a taxable benefit in his income each year.
3. Bernard must include the amount of the loan in his 2011 income.
4. Marc must include the amount of the loan in his 2011 income.
b. Sophie-Anne wants to acquire a 20% interest in Bigprofit Inc. She must pay $250,000 to the
corporation, which will issue her 10,000 voting and participating shares of the corporation,
representing 20% of the voting and participating shares of the corporation after the issue. Why
would Sophie-Anne want shares of a separate class from that of the current shareholders, who
had subscribed their shares for $10 each?
1. She wants the PUC of her shares to be $250,000.
2. She...
Attachments:
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THere four attachments. Assignment should be 2-3...
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Hello, Can you please check the attachment and see if you can do the project as following the instructions. Topic: A study on relationship between training and employee performance. ...
would you please follow the instructions in the file below the other attachment may help you to see how this case should be done thank you
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Commenter::, so it stands to reason that
IDropTarget is the interface that will be used.
From context, therefore, you should expect that the shell is going to
simulate a drop on your drop target. *g_ppr;
Of course our custom drop target needs a class ID:
const CLSID CLSID_Scratch = { ... };
I leave it to you to fill in the
CLSID structure
from the output of.
Remember that COM rules specify that the class factory itself
does not count as an outstanding COM object, so we don't use
the same
m_punkProcess trick that we did with
our drop target.
Instead, we just use a static object.RESULT hrRegister; DWORD dwRegisterCookie; { ProcessReference ref; g_ppr = &ref; hrRegister = CoRegisterClassObject(CLSID_Scratch, &s_scf, CLSCTX_LOCAL_SERVER, REGCLS_MULTIPLEUSE, &dwRegisterCookie);.)
After our message loop exits, we clear the
g_hwndChild
so
OpenFilesFromDataObject
knows that there is no
main window any more.
In real life, we'd have to create a new main window and restart
the message loop.\ so people will have
a clue what your server is for.::,
your code springs into action and extracts all the files in the
data object..
Nearly all of the work here was just managing the COM local server. The parts that had to do with the shell were actually quite small.
I have a few questions about a simulated drop (from both sides of the interface).
Is there a way for the drop target to detect a simulated drop (vs a real drop with the mouse)?
When the shell simulates a drop, does it call BOTH DragEnter and DragOver? What does it pass for key state, mouse point and effect?
If I have to simulate a drop (let's say I want to copy a file to an arbitrary IShellFolder), am I required to call DragEnter? What about DragOver? What mouse point to use?
It wouldn't be a very good simulation if there was!
In the spirit of Raymond, suck it and see! You could implement these interface members, and use OutputDebugString to check. That's my usual approach to these things.
As for you doing it... if you don't want other programs detecting you're a fake, it would be best to be as thorough as possible.
Iain.
And if I drop a file named "-Embedding"?
And if I want to do it in assembly?
Seriously, what drives such questions?
@Iain:
Let's say I'm writing a namespace extension. I want to detect when you drop a file at specific coordinates, or when you do a "paste" operation. For the first I have to use the mouse position, and for the second I have to ignore it. Would be nice to know which is the case.
I can argue that this is contrary to the "Raymond spirit". This way I will start depending on some undocumented behavior, possibly subject to change in the next version. Imagine if DragOver is being called. I do some processing there instead of in DragEnter/Drop. Well, Windows 11 may stop calling it and I'm screwed. If the actual sequence is documented, I can write my code correctly and future-proof it. Or maybe to create a shortcut, the shell is telling me that both Shift and Ctrl are pressed, instead of sending DROPEFFECT_LINK (for compatibility with some popular accounting software). Depending on such observed, but undocumented behavior is dangerous and should be avoided if there is an alternative.
I don't care if the program detects that I'm a fake (in fact it may be beneficial because of my first point). I want to paste a file (IDataObject) into a folder (IShellFolder with IDropTarget). There are multiple sequences I can use, but the only one that is guaranteed to work is the one that shell is using. That's because (I'm assuming) the drop target has been tested with the shell's fake drop. So the closer my behavior is to the shell, the better chances I have to not break anything.
>> And if I drop a file named "-Embedding"?
>
> Seriously, what drives such questions?
I'm not sure if this was an attempt at knocking the technique or a legitimate question - on Unix systems, filenames that look like switches is a problem that comes up often enough (especially because the shell expands wildcards, so the user might not even be aware there's such a file being passed on the command line) that programs often have a special switch ("--") to indicate that everything that follows is a filename even if it looks like a switch/option.
Great post Raymond. I'm just wondering about one thing, in this line
HDROP hdrop = reinterpret_cast<HDROP>(stgm.hGlobal);
shouldn't you use GlobalLock? Or does the shell guarantee its DataObject won't bother you with ancient relics like movable memory blocks? If so, is that true for all DataObjects one gets from the shell?
"[HKLMSoftwareMicrosoftWindowsCurrentVersionApp Pathsscratch.exe]
"
And if this is a per user app with its COM stuff in HKCU? AFAIK App Paths is HKLM only for some stupid reason, no UAC-less install for you!
I know this is not provided as 100% tested bug-free code, but for the sake of people who will undoubtedly copy and paste it, shouldn't there be a test for Query failure; in the code:
HRESULT hr = pdt->QueryInterface(riid, ppv);
pdt->Release();
?
I know that it's unlikely you will get queried about another interface not supported, but still, I think that Release() should be conditional.
Although I love articles about the deep stuff, I think articles like this are fine too once in a while, especially for people who are more susceptible to a hand on approach than to reading the docs (although they still should after they roughly understand what they're doing). Besides, I've ran into command line length limits in the past, so this may come in handy someday, and given that the COM stuff can be handled by templates or a framework, I think this is actually easier than the antiquated command line interface.
Seriously, it's 2010, console support should be dropped. If people really want a console they can code one up themselves. ;-)
That and the part you tell people sleep and wake now makes me giggle. :)
This sample should have been documented in the SDK 15 years ago. Why isn't it?
Amazing! I was just working on this very problem. Thanks Raymond.
Pierre B: *ppv, not pdt, is set to the result of QueryInterface. The caller to CreateInstance owns *ppv if it is set, not us. We own pdt, so need to release it before return.
Maybe it's nitpicking, but isn't OleInitialize() required instead of CoInitialize() when messing with IDataObject?
I'm dubious about the App Paths registry entry..
What if two programs in the system are named scratch.exe?
Wouldn't that registry entry redirect launching of all scratch.exe programs (target of file dropping) to launching the LocalServer32 entry instead?
(actually I tested a bit, and indeed it seems to be affecting file dropping on all scratch.exe)
You'd better choose carefully the name of your executable then..
Why is there no way for the executable itself (before executing) to tell the shell that it supports DropTarget? (for example, a special resource entry)
> [Um, the feature didn't exist 15 years ago. -Raymond]
You are probably lying, but let's pretend your not. Then replace 15 with 14 or "very long time ago".
You should invent and patent a markup language for articles like this. That would be a great tool to have in many technical works where the authors have only a foggy idea of who their target audience is*. Imagine an ebook transforming itself to your skillset. Brilliant.
* What you saying? That authors should figure out what "target audience" means? Don't be silly, that's HARD WORK!
"You are probably lying"? I guess blogs get the audience they deserve -- if the host self-diagnoses as having the social skills of a thermonuclear device, he'll get likewise readers.
@JM
Having the social skills of a thermonuclear device does not imply that one is an idiot. Raymond clearly is not, while yourself and sample are.
Nice, thanks!
Questions, please?
How do I know that I should use this cast: reinterpret_cast<HDROP>(stgm.hGlobal)?
Is it OK to use it if tymed is not TYMED_HGLOBAL? Or Is tymed guaranteed to be TYMED_HGLOBAL when IDataObject is coming from IDropTarget::Drop? Or does DragQueryFile know how to handle random hDrop? Or...?
[Exercise: Study the rules for IDataObject, then apply those rules to the situation here. -Raymond]
Can anyone play this game? The FORMATETC tells you how GetData should pass back the HDROP. You specified TYMED_HGLOBAL, so the HDROP was given back in a HGLOBAL. If you'd asked for something from GetData that can't fit in an HGLOBAL, it should return DV_E_FORMATETC. (In general, that is. I don't know why you'd be asking for anything larger in the case under discussion.)
Oh, and I forgot to say: The documentation for GetData says that you can bitwise or multiple TYMED_* types, and the callee can pick which type to give back. I didn't realize that before this, and it means that you would have to check what GetData returned in stgm.tymed. I wonder how many implementations of IDataObject don't handle that nuance?
Raymond, that little bit of knowledge alone makes the entire article worthwhile. Thanks!
@The_Assimilator: In my book, "you're lying" is about the worst thing you could say to a technical person, as it implies they're putting their ego before their profession. I'm an idiot for calling someone out for being rude to the author of the blog in an unnecessary and unlikely manner? I'm not sure how that works, but I'll choose to believe you just misunderstood what I wrote. Either that or your opinion is based on previous experience, in which case due apologies for whatever I did in the past...
In any case, let's drop it, I already regret commenting at all. | https://blogs.msdn.microsoft.com/oldnewthing/20100503-00/?p=14183 | CC-MAIN-2017-30 | refinedweb | 1,688 | 73.47 |
Query against an Action?AdamWilden Dec 6, 2011 3:52 AM
Hi Folks,
Still being pestered by the Service Desk for more stats (their manager went on a Course ) so I've started pushing as much as possible onto themselves via WebDesk, which allows them to knock up quite a lot of their queries themselves.
But I'm stuck on quite an obvious one (at least I think it's obvious but I've only ever run stats against collections).
Our users don't categorise when logging calls, our Service Desk does this. They select an Incident category and then click on a winodwless action (called Categoirised).
This is ususally the frst action that SD staff carry out so they understandably need to know the average time from the call being first logged to being categorised.
(Actually they want to know this for each category, and almost certainly they'll want the stats to sing and dance... ).
Any idea how I can access the timestamp for when the call was categorised? Or do I need to hit the audit table? (and if so - help!).
Many thanks.
Adam.
1. Re: Query against an Action?Julian Wigman Dec 6, 2011 7:45 AM (in response to AdamWilden)
Adam,
My suggestion wouldn't help for existing Incidents but what I was thinking would be to have a datetime attribute on the Incident BO that is set from an automatic action following your "categorised" windowless action in the process. In the field on the automatic action use a BOO calculation and set to now().
This way it stores the date time it was categorised for easy reporting later.
Regards
Julian
2. Re: Query against an Action?karenpeacock Dec 6, 2011 8:24 AM (in response to Julian Wigman)
Hi
If you would like to extract data from the Audit Trail here is some information about this:
Best wishes
Karen
3. Re: Query against an Action?AdamWilden Dec 6, 2011 11:42 AM (in response to Julian Wigman)
Thanks Guys,
I think the auto action is probably simplest but having said that I can't get it working!
I've added a DateTime attribute to the Incident BO but think I must be going wrong somewhere on the Process side of things.
I've created an Automatic Action immediately after my "Categorised" windowless action.
Not sure where to add the calc though - If I double click the new action, my Incident window opens and I can add the calculation to my new attribute that way which is hwat I've done.
To test I open a new Incident - the new attribute automatically fills with the current date/time. I wait a few minutes and click the Categorise action but the new date field does not update. I can see that the process is flowing correctly
The calculation is simple:
import System
static def GetAttributeValue(Incident):
Value = DateTime.Now
return Value
Not sure where I'm going wrong.
Thanks.
4. Re: Query against an Action?AdamWilden Dec 7, 2011 4:35 AM (in response to AdamWilden)
Hi Guys,
They now also want to be able to report on who originally categorised the call as well as teh timestamp so I think it's probably easiest to create a new collection to record the time and current user etc.
I should be able to cope with this but many thanks for your suggestions and advice.
Cheers - Adam.
5. Re: Query against an Action?AdamWilden Oct 15, 2013 8:54 AM (in response to AdamWilden)
Hi Guys,
This one has come back again so back looking for advice
Just need some way of creating a query within WebDesk to get business time between creation date and our "Categorised" status which follows a windowless action.
The actual calc will be simple:
TimePeriod = Incident.GetBusinessTime(Incident.CreationDate, *Categorised-Date*)
return Int32.Parse(String.Format("{0}", Math.Floor(TimePeriod.TotalMinutes))) + 1
The solution doesn't need to unclude existing calls so can be something new...
Julian's suggestion seemed sensible but I didn't manage to get it working before.
Any bright ideas?
Cheers - A. | https://community.ivanti.com/thread/17493 | CC-MAIN-2018-34 | refinedweb | 684 | 61.87 |
Apache Software Foundation, new devices added by a Device File Pack are guaranteed to work in these versions of the IDE
Now Evaluate AVR and SAM MCUs with MPLAB X IDE v5.15
Want to evaluate Microchip’s AVR and SAM microcontrollers, but don’t have time to learn a new ecosystem? Now users of the MPLAB ecosystem can evaluate Microchip’s 8-bit AVR and 32-bit SAM devices through their favorite development tools with the release of MPLAB X IDE version 5.15. In addition, this version of our popular IDE fully supports all Microchip PIC and dsPIC microcontrollers as it always has. MPLAB X IDE version 5.15 gives support for:
- Most AVR families
- SAM E70/S70/V71, SAM D21/C21SAM, SAM E5X/D5X MCU microcontroller families and more
To see if your AVR or SAM device is supported, download MPLAB X IDE version 5.15 and see the New Project dialogue box
In addition to IDE support for AVR and SAM devices, debugging and compiling can be completed for AVR and SAM devices with this list of development tools:
- Get your AVR code off to a head-start with MPLAB Code Configurator
- The high performance MPLAB ICD 4 in-circuit debugger and programmer (SAM support only at this time)
- The ever popular MPLAB PICkit™ 4 in-circuit debugger and programmer
- The affordable MPLAB Snap in-circuit debugger
- The MPLAB XC8 C Compiler for AVR devices
- The MPLAB XC32++ C Compiler for SAM devices
- The AVR GCC Compiler for AVR devices
- The Arm® GCC compiler, for devices with the Arm architecture (SAM)
To view a complete listing of support, see the MPLAB X IDE Device Support List found in the documentation tab below.
Get started with your evaluation of AVR and SAM microcontrollers by either starting a project in Atmel START and importing it into MPLAB X IDE, or start right away inside MPLAB X IDE.
mplab x Screenshots:
).
In order to navigate your code or understand a colleague’s code in addition to documentation the Call Graph provides a static call tree of all functions called from other functions. It can also be exported to a Portable Network Graphics (PNG) image.
A single project can now build the same set of source files in many different ways. Each “configuration” has its own compiler options, compiler, hardware tool, and target device.
.
Using the CTRL key and mouse over a function, variable, macro, or include statement allows you to view its declaration. Clicking on the hyperlink will take you right to the source of declaration. Alternatively, you can right click on it and choose Navigate → Go to Declaration from the context menu to jump to its declaration.
Within the IDE there are many existing code templates that can be accessed using a couple of letters then tab (or specified key). You can create your own templates, (even live templates) such that when you enter values into the template area, other areas of code are also populated. For example the ‘func’ active code template shown here. As you enter parameters in the comments, to document the function, the real function is also populated. The developer can select from a template file whenever he creates a new file to add source code to.
Unsure of what changes you made to the software since the last version control update? Use the Local History utility to visualize changes made. Very useful for backtracking
Either an individual or a company can set up a code format standard to be used within the editor. Just select the file to format the code in and menu Source>/Format to reapply the template to your source code.
The Tasks operation, automatically scans your code and lists commented lines containing words such as “TODO” or “FIXME”, (the words can be customized under options). Tasks provide a convenient way to keep track of important items you feel need addressing.
Macros are incredible useful but sometimes they can have unexpected values if they are conditionally defined. This window allows you to see what the compiler will consume after the preprocessor is done. With the expansion view you see exactly what value they expand to. Also, blocks of code not to be compiled are omitted in the view. Also, in the editor window, MPLAB X shows you all the #ifdef/#endif blocks. It uses the comment color (grey by default) to show you sections that will not be included.
You can change any memory view to look at any type of memory. Formats for those views are also selectable from the dropdowns. This allows a quick view change without going thru the menus.
Need one place to summarize you project? For convenience there is a single window that gathers all the relevant project information and its environment. Device name, Debug Tool, Language Build Tool, and Connection state are presented. The Memory section shows Total, Used and Reserved by Debug Tool for RAM and Flash memory. Checksum and Breakpoint (silicon resource) status is also shown. The Debug tool provides additional status for Device ID, firmware versions and voltages.
With this feature you’ll never have to worry about which firmware version you were using. MPLAB will remember it and automatically restore it to the correct version when you connect to the debugger.
Just set it and forget it.
.
Use the Variables and Call Stack windows together to browse local variable history of each function in the call chain. Hint: The Variables window is docked at the bottom of the output pane to allow simultaneous interaction and display of the local variables with the selected function in the Call Stack window
.
mplab x ide download link:
mplab x quick video overview:
Source : MPLAB X IDE download
Current Project / Post can also be found using:
- mplab x ide
- mplab x ide download
- Mplab x
- mplab x download | https://pic-microcontroller.com/mplab-x-ide-download/ | CC-MAIN-2021-39 | refinedweb | 974 | 59.74 |
Creating C# objects at run-time is easy. Just call the Activator.CreateInstance method.
For example, to create the same type of object as the current object using its default constructor:
Activator.CreateInstance( this.GetType() );
To better illustrate how this is done, below is a complete console program you can run. In this simple example, there is a Shape base class, and some derived shapes such as Circle and Square. The object construction code is used in the Shape’s “New” method, which could be used by a Shape factory class (not shown).
By default, the “New” method uses “this.GetType()” to construct the same type of Shape as the current object. This eliminates the need for each derived class to implement the “New” method. But if desired, shapes can override the “New” method to alter the construction process somehow. In this example, the SmallSquare class overrides the “New” method to ensure that no squares larger than 20×20 are created.
using System; using System.Drawing; namespace ConstructorThroughReflection { class Program { static void Main( string[] args ) { Circle c1 = new Circle( new Size( 2, 3 ) ); Circle c2 = c1.New( new Size( 10, 5 ) ) as Circle; Square s1 = new Square( new Size( 3, 4 ) ); SmallSquare ss1 = new SmallSquare( new Size( 7, 1 ) ); SmallSquare ss2 = new SmallSquare( new Size( 100, 100 ) ); Console.ReadLine(); } } abstract public class Shape { public Shape() {} public Shape( Size size ) { Console.WriteLine( "Shape={0}, Size={1}", this.GetType().Name, size ); this.Size = size; } public Size Size; virtual public Shape New( Size size ) { return Activator.CreateInstance( this.GetType(), size ) as Shape; } } public class Circle : Shape { public Circle() {} public Circle( Size size ) : base( size ) {} } public class Square : Shape { public Square() { } public Square( Size size ) : base( size ) { } } public class SmallSquare : Square { public SmallSquare() { } public SmallSquare( Size size ) : base( size ) { } override public Shape New( Size size ) { const int c_MaxSize = 20; int width = size.Width; int height = size.Height; if (width > c_MaxSize) width = c_MaxSize; if (height > c_MaxSize) height = c_MaxSize; size = new Size( width, height ); return new SmallSquare( size ); } } }
As you might expect, the console output from this program is:
Shape=Circle, Size={Width=2, Height=3}
Shape=Circle, Size={Width=10, Height=5}
Shape=Square, Size={Width=3, Height=4}
Shape=SmallSquare, Size={Width=7, Height=1}
Shape=SmallSquare, Size={Width=20, Height=20} | http://www.csharp411.com/construct-csharp-objects-at-runtime/ | CC-MAIN-2017-22 | refinedweb | 378 | 63.59 |
On Tue, Oct 16, 2012 at 9:37 PM, AUGER Cédric <sedrikov at gmail.com> wrote: > As I said, from the mathematical point of view, join (often noted μ in > category theory) is the (natural) transformation which with return (η > that I may have erroneously written ε in some previous mail) defines a > monad (and requires some additionnal law). Auger: Your emails keep invoking "the mathematical point of view" as if it were something unique and universal.. I don't understand your dogmatism about return and join being canonically the best monad definition in all possible mathematics. That's truly a quantifier that beggars imagination. -- Kim-Ee On Tue, Oct 16, 2012 at 9:37 PM,: <> | http://www.haskell.org/pipermail/haskell-cafe/2012-October/104083.html | CC-MAIN-2014-35 | refinedweb | 116 | 52.9 |
Hello readers, Ned here again. In the previous two blog posts I discussed planning for DFSR server replacements and how to ensure you are properly pre-seeding data. Now I will show how to replace servers in an existing Replication Group using the N+1 Method to minimize interruption.
Make sure you review the first two blog posts before you continue:
- Replacing DFSR Member Hardware or OS (Part 1: Planning)
- Replacing DFSR Member Hardware or OS (Part 2: Pre-seeding)
Background
As mentioned previously, the “N+1” method entails adding a new replacement server in a one-to-one partnership with the server being replaced. That new computer may be using local fixed storage (likely for a branch file server) or using SAN-attached storage (likely for a hub file server). Because replication is performed to the replacement server – preferably with pre-seeded data – the interruption to existing replication is minimal and there is no period where replication is fully halted. This reduces risk as there is no single point of failure for end users, and backups can continue unmolested in the hub site.
The main downside is cost and capacity. For each N+1 operation you need an equal amount of storage available to the new computer, at least until the migration is complete. It also means that you need an extra server available for the operation on each previous node (if doing a hardware refresh this is not an issue, naturally).
Because a new server is being added for each old server in N+1, that will run simultaneously until the old server is decommissioned.
- Enough storage for each replacement server to hold as much data as the old server.
- If replacing a server with a cluster, two or more replacement servers will be required (this is typically only done on the hub servers).
Phase 1 – Adding the new server. Bring the new DFSR server online.
3. Optional but recommended: Pre-seed the new server with existing data from the hub.
Note: for pre-seeding techniques, see Replacing DFSR Member Hardware or OS (Part 2: Pre-seeding)
4. Add the new server as a new member of the first replication group.
Note: For steps on using DFSR clusters, reference:
- Deploying DFS Replication on a Windows Failover Cluster – Part I
- Deploying DFS Replication on a Windows Failover Cluster – Part II
- Deploying DFS Replication on a Windows Failover Cluster – Part III
5. Select the server being replaced as the only replication partner with the new server. Do not select any other servers.
6. Create (or select, if pre-seeded) the new replicated folder path on the replacement server.
Note: Optionally, you can make this a Read-Only replicated folder if running Windows Server 2008 R2. Make sure you understand the RO requirements and limitation by reviewing:
7. Complete the setup. Allow AD replication to converge (or force it with REPADMIN.EXE /SYNCALL). Allow DFSR polling to discover the new configuration (or force it with DFSRDIAG.EXE POLLAD).
8. At this point, the new server is replicating only with the old server being replaced.
9. When done, the new server will log a 4104 event. If pre-seeding was done correctly then there will be next to no 4412 conflict events (unless the environment is completely static there are likely to be some 4412’s, as users will continue to edit data normally).
10. Repeat for any other Replication Groups or Replicated folders configured on the old server, until the new server is a configured identically and has all data.
Phase 2 – Recreate the replication topology
1. Select the Replication Group and create a “New Topology”.
2. Select a hub and spoke topology.
Note: You can use a full mesh topology with customization if using a more complex environment.
3. Make the new replacement server the new hub. The old server will act as a “spoke” temporarily until it is decommissioned; this allows for it to continue replicating any last minute user changes.
4. Force AD replication and DFSR polling again. Verify that all three servers are replicating correctly by creating a propagation test file using DFSRDIAG.EXE PropagationTest or DFSMGMT.MSC’s propagation test.
5. Create folder shares on the replacement server to match the old share names and data paths.
6. Repeat these steps above for any other RG’s/RF’s that are being replaced on these servers.
Phase 3 – Removing the old server
Note: this phase is the only one that potentially affects user file access. It should be done off hours in a change control window in order to minimize user disruption. In a reliably connected network environment with an administrator that is comfortable using REPADMIN and DFSRDIAG to speed up configuration convergence, the entire outage can usually be kept under 5 minutes.
1. Stop further user access to the old file server by removing the old shares.
Note: Stopping the Server service with command NET STOP LANMANSERVER will also temporarily prevent access to shares.
2. Remove the old server from DFSR replication by deleting the Member within all replication groups. This is done on the Membership tab by right-clicking the old server and selecting “Delete”.
3. Wait for the DFSR 4010 event(s) to appear for all previous RG memberships on that server before continuing.
4. At this point the old server is no longer allowing user data or replicating files. Rename the old server so that no accidental access can occur further. If part of DFS Namespace link targeting, remove it from the namespace as well..
6. Force AD replication and DFSR polling. Validate that the servers correctly see the name change.
7..
8. Replication can be confirmed as continuing to work after the rename as well.
9. The process is complete.
Final Notes
As you can now see the steps to perform an N+1 migration operation are straightforward no matter if replacing a hub, branch, or all servers. Use of DFS Namespaces makes this more transparent to users. The actual outage time of N+1 is theoretically zero if not renaming servers and performing the operation off hours when users are not actively accessing data. Replication to the main office for never stops, so centralized backups can continue during the migration process.
All of these factors make N+1 the recommended DFSR node replacement strategy.
Series Index
– Ned “+1” Pyle | https://blogs.technet.microsoft.com/askds/2010/09/08/replacing-dfsr-member-hardware-or-os-part-3-n1-method/ | CC-MAIN-2018-43 | refinedweb | 1,059 | 55.13 |
As author for ebook for epub, all carriage return (or enter) show as a question marks. Why?
I doubt you're using Authorware to make an eBook....? What software tool are you using?
Best to post in that forum....likely InDesign?
That said, it probably has something to do with the font you're using (embed it with the project?) or the encoding (ASCII vs UTF-8 vs Unicode) you have setup...maybe.
Erik
Hi Erik - I used Abobe Digital Editions.
Pam
Erik
I am on a mac and want to check my .ePub file. Opened manuscript in Digital Editions. The jacket did not show at all, some page breaks (or might be section breaks) are ignored and ALL of the carriage returns (or ENTERS) are question marks. Presumably Digital Editions is the right place, and not Authorware which I have only just found?
Pam | http://forums.adobe.com/thread/973600 | CC-MAIN-2014-15 | refinedweb | 145 | 75.5 |
Source:NetHack 3.6.0/src/write.c
From NetHackWiki
(Redirected from Source:Write.c)
Below is the full text to write.c from the source code of NetHack 3.6.0. To link to a particular line, write [[Source:NetHack 3.6.0/src/write.c.
Top of file
/* NetHack 3.6 write.c $NHDT-Date: 1446078770 2015/10/29 00:32:50 $ $NHDT-Branch: master $:$NHDT-Revision: 1.16 $ */
/* NetHack may be freely redistributed. See license for details. */
#include "hack.h"
STATIC_DCL int FDECL(cost, (struct obj *));
STATIC_DCL boolean FDECL(label_known, (int, struct obj *));
STATIC_DCL char *FDECL(new_book_description, (int, char *));
cost
/*
* returns basecost of a scroll or a spellbook
*/
STATIC_OVL int
cost(otmp)
register struct obj *otmp;
{
if (otmp->oclass == SPBOOK_CLASS)
return (10 * objects[otmp->otyp].oc_level);
switch (otmp->otyp) {
#ifdef MAIL
case SCR_MAIL:
return 2;
#endif
case SCR_LIGHT:
case SCR_GOLD_DETECTION:
case SCR_FOOD_DETECTION:
case SCR_MAGIC_MAPPING:
case SCR_AMNESIA:
case SCR_FIRE:
case SCR_EARTH:
return 8;
case SCR_DESTROY_ARMOR:
case SCR_CREATE_MONSTER:
case SCR_PUNISHMENT:
return 10;
case SCR_CONFUSE_MONSTER:
return 12;
case SCR_IDENTIFY:
return 14;
case SCR_ENCHANT_ARMOR:
case SCR_REMOVE_CURSE:
case SCR_ENCHANT_WEAPON:
case SCR_CHARGING:
return 16;
case SCR_SCARE_MONSTER:
case SCR_STINKING_CLOUD:
case SCR_TAMING:
case SCR_TELEPORTATION:
return 20;
case SCR_GENOCIDE:
return 30;
case SCR_BLANK_PAPER:
default:
impossible("You can't write such a weird scroll!");
}
return 1000;
}
label_known
/* decide whether the hero knowns a particular scroll's label;
unfortunately, we can't track things are haven't been added to
the discoveries list and aren't present in current inventory,
so some scrolls with ought to yield True will end up False */
STATIC_OVL boolean
label_known(scrolltype, objlist)
int scrolltype;
struct obj *objlist;
{
struct obj *otmp;
/* only scrolls */
if (objects[scrolltype].oc_class != SCROLL_CLASS)
return FALSE;
/* type known implies full discovery; otherwise,
user-assigned name implies partial discovery */
if (objects[scrolltype].oc_name_known || objects[scrolltype].oc_uname)
return TRUE;
/* check inventory, including carried containers with known contents */
for (otmp = objlist; otmp; otmp = otmp->nobj) {
if (otmp->otyp == scrolltype && otmp->dknown)
return TRUE;
if (Has_contents(otmp) && otmp->cknown
&& label_known(scrolltype, otmp->cobj))
return TRUE;
}
/* not found */
return FALSE;
}
dowrite
static NEARDATA const char write_on[] = { SCROLL_CLASS, SPBOOK_CLASS, 0 };
/* write -- applying a magic marker */
int
dowrite(pen)
register struct obj *pen;
{
register struct obj *paper;
char namebuf[BUFSZ], *nm, *bp;
register struct obj *new_obj;
int basecost, actualcost;
int curseval;
char qbuf[QBUFSZ];
int first, last, i, deferred, deferralchance;
boolean by_descr = FALSE;
const char *typeword;
if (nohands(youmonst.data)) {
You("need hands to be able to write!");
return 0;
} else if (Glib) {
pline("%s from your %s.", Tobjnam(pen, "slip"),
makeplural(body_part(FINGER)));
dropx(pen);
return 1;
}
/* get paper to write on */
paper = getobj(write_on, "write on");
if (!paper)
return 0;
/* can't write on a novel (unless/until it's been converted into a blank
spellbook), but we want messages saying so to avoid "spellbook" */
typeword = (paper->otyp == SPE_NOVEL)
? "book"
: (paper->oclass == SPBOOK_CLASS)
? "spellbook"
: "scroll";
if (Blind) {
if (!paper->dknown) {
You("don't know if that %s is blank or not.", typeword);
return 1;
} else if (paper->oclass == SPBOOK_CLASS) {
/* can't write a magic book while blind */
pline("%s can't create braille text.",
upstart(ysimple_name(pen)));
return 1;
}
}
paper->dknown = 1;
if (paper->otyp != SCR_BLANK_PAPER && paper->otyp != SPE_BLANK_PAPER) {
pline("That %s is not blank!", typeword);
exercise(A_WIS, FALSE);
return 1;
}
/* what to write */
Sprintf(qbuf, "What type of %s do you want to write?", typeword);
getlin(qbuf, namebuf);
(void) mungspaces(namebuf); /* remove any excess whitespace */
if (namebuf[0] == '\033' || !namebuf[0])
return 1;
nm = namebuf;
if (!strncmpi(nm, "scroll ", 7))
nm += 7;
else if (!strncmpi(nm, "spellbook ", 10))
nm += 10;
if (!strncmpi(nm, "of ", 3))
nm += 3;
if ((bp = strstri(nm, " armour")) != 0) {
(void) strncpy(bp, " armor ", 7); /* won't add '\0' */
(void) mungspaces(bp + 1); /* remove the extra space */
}
deferred = 0; /* not any scroll or book */
deferralchance = 0; /* incremented for each oc_uname match */
first = bases[(int) paper->oclass];
last = bases[(int) paper->oclass + 1] - 1;
for (i = first; i <= last; i++) {
/* extra shufflable descr not representing a real object */
if (!OBJ_NAME(objects[i]))
continue;
if (!strcmpi(OBJ_NAME(objects[i]), nm))
goto found;
if (!strcmpi(OBJ_DESCR(objects[i]), nm)) {
by_descr = TRUE;
goto found;
}
/* user-assigned name might match real name of a later
entry, so we don't simply use first match with it;
also, player might assign same name multiple times
and if so, we choose one of those matches randomly */
if (objects[i].oc_uname && !strcmpi(objects[i].oc_uname, nm)
/*
* First match: chance incremented to 1,
* !rn2(1) is 1, we remember i;
* second match: chance incremented to 2,
* !rn2(2) has 1/2 chance to replace i;
* third match: chance incremented to 3,
* !rn2(3) has 1/3 chance to replace i
* and 2/3 chance to keep previous 50:50
* choice; so on for higher match counts.
*/
&& !rn2(++deferralchance))
deferred = i;
}
/* writing by user-assigned name is same as by description:
fails for books, works for scrolls (having an assigned
type name guarantees presence on discoveries list) */
if (deferred) {
i = deferred;
by_descr = TRUE;
goto found;
}
There("is no such %s!", typeword);
return 1;
found:
if (i == SCR_BLANK_PAPER || i == SPE_BLANK_PAPER) {
You_cant("write that!");
pline("It's obscene!");
return 1;
} else if (i == SPE_BOOK_OF_THE_DEAD) {
pline("No mere dungeon adventurer could write that.");
return 1;
} else if (by_descr && paper->oclass == SPBOOK_CLASS
&& !objects[i].oc_name_known) {
/* can't write unknown spellbooks by description */
pline("Unfortunately you don't have enough information to go on.");
return 1;
}
/* KMH, conduct */
u.uconduct.literate++;
new_obj = mksobj(i, FALSE, FALSE);
new_obj->bknown = (paper->bknown && pen->bknown);
/* shk imposes a flat rate per use, not based on actual charges used */
check_unpaid(pen);
/* see if there's enough ink */
basecost = cost(new_obj);
if (pen->spe < basecost / 2) {
Your("marker is too dry to write that!");
obfree(new_obj, (struct obj *) 0);
return 1;
}
/* we're really going to write now, so calculate cost
*/
actualcost = rn1(basecost / 2, basecost / 2);
curseval = bcsign(pen) + bcsign(paper);
exercise(A_WIS, TRUE);
/* dry out marker */
if (pen->spe < actualcost) {
pen->spe = 0;
Your("marker dries out!");
/* scrolls disappear, spellbooks don't */
if (paper->oclass == SPBOOK_CLASS) {
pline_The("spellbook is left unfinished and your writing fades.");
update_inventory(); /* pen charges */
} else {
pline_The("scroll is now useless and disappears!");
useup(paper);
}
obfree(new_obj, (struct obj *) 0);
return 1;
}
pen->spe -= actualcost;
/*
* Writing by name requires that the hero knows the scroll or
* book type. One has previously been read (and its effect
* was evident) or been ID'd via scroll/spell/throne and it
* will be on the discoveries list.
* (Previous versions allowed scrolls and books to be written
* by type name if they were on the discoveries list via being
* given a user-assigned name, even though doing the latter
* doesn't--and shouldn't--make the actual type become known.)
*
* Writing by description requires that the hero knows the
* description (a scroll's label, that is, since books by_descr
* are rejected above). BUG: We can only do this for known
* scrolls and for the case where the player has assigned a
* name to put it onto the discoveries list; we lack a way to
* track other scrolls which have been seen closely enough to
* read the label without then being ID'd or named. The only
* exception is for currently carried inventory, where we can
* check for one [with its dknown bit set] of the same type.
*
* Normal requirements can be overridden if hero is Lucky.
*/
/* if known, then either by-name or by-descr works */
if (!objects[new_obj->otyp].oc_name_known
/* else if named, then only by-descr works */
&& !(by_descr && label_known(new_obj->otyp, invent))
/* and Luck might override after both checks have failed */
&& rnl(Role_if(PM_WIZARD) ? 5 : 15)) {
You("%s to write that.", by_descr ? "fail" : "don't know how");
/* scrolls disappear, spellbooks don't */
if (paper->oclass == SPBOOK_CLASS) {
You(
"write in your best handwriting: \"My Diary\", but it quickly fades.");
update_inventory(); /* pen charges */
} else {
if (by_descr) {
Strcpy(namebuf, OBJ_DESCR(objects[new_obj->otyp]));
wipeout_text(namebuf, (6 + MAXULEV - u.ulevel) / 6, 0);
} else
Sprintf(namebuf, "%s was here!", plname);
You("write \"%s\" and the scroll disappears.", namebuf);
useup(paper);
}
obfree(new_obj, (struct obj *) 0);
return 1;
}
/* can write scrolls when blind, but requires luck too;
attempts to write books when blind are caught above */
if (Blind && rnl(3)) {
/* writing while blind usually fails regardless of
whether the target scroll is known; even if we
have passed the write-an-unknown scroll test
above we can still fail this one, so it's doubly
hard to write an unknown scroll while blind */
You("fail to write the scroll correctly and it disappears.");
useup(paper);
obfree(new_obj, (struct obj *) 0);
return 1;
}
/* useup old scroll / spellbook */
useup(paper);
/* success */
if (new_obj->oclass == SPBOOK_CLASS) {
/* acknowledge the change in the object's description... */
pline_The("spellbook warps strangely, then turns %s.",
new_book_description(new_obj->otyp, namebuf));
}
new_obj->blessed = (curseval > 0);
new_obj->cursed = (curseval < 0);
#ifdef MAIL
if (new_obj->otyp == SCR_MAIL)
new_obj->spe = 1;
#endif
new_obj =
hold_another_object(new_obj, "Oops! %s out of your grasp!",
The(aobjnam(new_obj, "slip")), (const char *) 0);
return 1;
}
new_book_description
/* most book descriptions refer to cover appearance, so we can issue a
message for converting a plain book into one of those with something
like "the spellbook turns red" or "the spellbook turns ragged";
but some descriptions refer to composition and "the book turns vellum"
looks funny, so we want to insert "into " prior to such descriptions;
even that's rather iffy, indicating that such descriptions probably
ought to be eliminated (especially "cloth"!) */
STATIC_OVL char *
new_book_description(booktype, outbuf)
int booktype;
char *outbuf;
{
/* subset of description strings from objects.c; if it grows
much, we may need to add a new flag field to objects[] instead */
static const char *const compositions[] = {
"parchment",
"vellum",
"cloth",
#if 0
"canvas", "hardcover", /* not used */
"papyrus", /* not applicable--can't be produced via writing */
#endif /*0*/
0
};
const char *descr, *const *comp_p;
descr = OBJ_DESCR(objects[booktype]);
for (comp_p = compositions; *comp_p; ++comp_p)
if (!strcmpi(descr, *comp_p))
break;
Sprintf(outbuf, "%s%s", *comp_p ? "into " : "", descr);
return outbuf;
}
/*write.c*/ | https://nethackwiki.com/wiki/Source:Write.c | CC-MAIN-2017-34 | refinedweb | 1,646 | 51.58 |
LinuxQuestions.org
(
/questions/
)
-
Linux - Networking
(
)
- -
Unknown Host <Linuxmachinename> / Unable to ping by host name
(
)
nishi_k_79
10-28-2003 11:49 PM
Unknown Host <Linuxmachinename> / Unable to ping by host name
Hi,
We have windows 98,win2000 and Linux connected in a workgroup. I can ping my Linux machine by ip address but not by host name.In the Windows Explorer of win 98 if i type the ip adrress i can browse my Linux shares..and consequently if i type machinename i am able to browse my shares.This is as far as Win98 behaviour is concerned . But i am unable to browse by ipaddr in Window 2000.
Net View \\<Linuxmachine> in Windows 98 shows all shares whereas not the same in Win 2000.
Also, i am still not able to ping by name from either machines.
Any help greatly appreciated.I am already at my wits end !!
Also, Firewall is turned off.
pnh73
10-29-2003 01:52 AM
To ping by host name you need to have some kind of hostname resolution such as a DNS server to find out what the IP of that host is. Can you browse 98 shares from 2000?
I am not sure about your Windows 2000 problem as I have never had experience of Windows 2000. It is likely that 2000 doesnt have the necessary protocol installed. I think that the necessarry protocol is called "Microsoft File and Print Sharing" or something. Look in the network settings in Windows 98 and make sure you have the same installed on win 2000.
HTH
nishi_k_79
10-30-2003 04:34 AM
Hi,
I have my DNS Resolution enabled.Also, i can browse 98 shares from 2000 and 2000 has Microsoft File and Print Sharing enabled.
Also, I can access the Internet from my Linux Machine (Gateway)..however if i try to access it from other machines setting Linux m/c as gateway ..i cannot access the Internet.
Running Traceroute shows following behaviour.
1) On Linux :
In 1 hop it reaches 192.168.0.X(external gateway)
2) On Win machine:
1st hop -> Linux machine
2nd hop -> Request Timed Out !!
zaphodiv
10-31-2003 10:30 AM
Internet domain names are a different namespace to NETBIOS names, or should be, windows sometimes mixes the two in troublesome ways.
>To ping by host name you need to have some kind of hostname resolution such as a DNS server to find out
>what the IP of that host is.
>I have my DNS Resolution enabled.
I expect you are sending DNS querys to your ISP's resolving nameservers. Your ISP's server knows nothing of the machines on your home LAN.
The easiest way is to add a line in the host file on every machine that maps your desired hostname to an ip address.
on 2k/XP the host files is somthing like
c:\winnt\system32\drivers\etc\hosts
pnh73
11-01-2003 01:24 PM
The same can be done in Linux with /etc/hosts
Format is:
domain this.is.the.ip
All times are GMT -5. The time now is
10:19 AM
. | http://www.linuxquestions.org/questions/linux-networking-3/unknown-host-linuxmachinename-unable-to-ping-by-host-name-109772-print/ | CC-MAIN-2014-35 | refinedweb | 519 | 74.08 |
vdiff 2.4.0
Efficiently manage the differences between two files using vim.
Opens two files in vimdiff and provides single-stroke key mappings to make moving differences between two files efficient. Up to two additional files may be opened at the same time, but these are generally used for reference purposes.
Usage
vdiff [options] <file1> <file2> [<file3> [<file4>]]
Options
Relevant Key Mappings
Defaults
Defaults will be read from ~/.config/vdiff/config if it exists. This is a Python file that is evaluated to determine the value of three variables: vimdiff, gvimdiff, and gui. The first two are the strings used to invoke vimdiff and gvimdiff. The third is a boolean that indicates which should be the default. If gui is true, gvimdiff is used by default, otherwise vimdiff is the default. An example file might contain:
vimdiff = 'gvimdiff -v' gvimdiff = 'gvimdiff -f' gui = True
These values also happen to be the default defaults.
As a Package
You can also use vdiff in your own Python programs. To do so, you would do something like the following:
from inform import Error from vdiff import Vdiff vdiff = Vdiff(lfile, rfile) try: vdiff.edit() except KeyboardInterrupt: vdiff.cleanup() except Error as err: err.report()
Installation
Runs only on Unix systems. Requires Python 3.5 or later. Install by running ‘./install’ or ‘pip3 install vdiff’.
- Author: Ken Kundert
- Download URL:
- Keywords: vim,diff
- License: GPLv3+
- Categories
- Development Status :: 5 - Production/Stable
- Environment :: Console
- Intended Audience :: End Users/Desktop
- License :: OSI Approved :: GNU General Public License v3 or later (GPLv3+)
- Natural Language :: English
- Operating System :: POSIX :: Linux
- Programming Language :: Python :: 3.5
- Topic :: Utilities
- Package Index Owner: kenkundert
- DOAP record: vdiff-2.4.0.xml | https://pypi.python.org/pypi/vdiff/ | CC-MAIN-2017-51 | refinedweb | 281 | 50.33 |
Here is the code:
#include <iostream> using namespace std; void intro() { cout << "Welcome to CASTLE" << endl; cin.get(); cout << "A midevil RPG. Source is avaliable at thetrh51.blogspot.com along with documentation." << endl; cout << "All the commands, descriptions are avaliable at thetrh51.blogspot.com"; cin.get(); } int menu() { cout << "please visit thetrh51.blogspot.com for the source and info on the program including commands." << endl; int dec; cout << "1. save" << endl << "2. load" << endl << "3. contiue" << endl << "4. exit" << endl; cout << "Please select from above: "; cin >> dec; } void save() { } char getinfo() { cout << "Please enter your characters name [max 12 characters]: "; char playername[12]; cin >> playername; cout << endl << "OK " << playername << " how old do you want to be? [20-90]: "; int playerage; cin >> playerage; cout << endl << playername << " what is your difficulty going to be? [1-10]: "; int difficulty; cin >> difficulty; cout << endl << "Here is what you entered" << endl; cout << "Your name is " << playername << endl; cout << "Your age is " << playerage << endl; cout << "Your dificulty level is " << difficulty << endl; } int main() { int score; licence(); intro(); getinfo(); char command[12]; cout << "command: "; cin >> command; if (command == "menu") { menu(); cin.get(); } else if (command == "help") { cout << "visit thetrh51.blogspot.com for more info."; cin.get(); } else if (command == "exit") { return 1; } }
PS. If any one wants to help me with the save file it would be appreciated.
Attached File(s)
castle.txt (1.48K)
Number of downloads: 40 | http://www.dreamincode.net/forums/topic/280326-having-problems-with-c-code-else-if-statements-i-think/ | CC-MAIN-2016-07 | refinedweb | 232 | 77.53 |
Comment on Tutorial - How to Send SMS using Java Program (full code sample included) By Emiley J.
Comment Added by : Yanosh
Comment Added at : 2009-07-16 05:19:38
Comment on Tutorial : How to Send SMS using Java Program (full code sample included) By Emiley J.
Hi,
I tried use this code:
package sms;
public class Main {
public static void main(String[] args) {
// TODO code application logic here
SMSClient sc = new SMSClient(0);
sc.sendMessage("505050702","Hello");
}
}
And I have error when tried to run in NetBeans and in console:
Exception in thread "main" java.lang.NoClassDefFoundError: sms/SMSClient (wrong name: SMSClient)
Pls. help
Yan. No it won't work...
View Tutorial By: Arin at 2012-08-17 09:59:40
2. To execute program you need comm.jar that is not a
View Tutorial By: shreyansh shah at 2012-07-11 05:57:42
3. Unable to connect with https:\\localhost:8443 but
View Tutorial By: gopala krishna at 2012-04-02 07:17:09
4. Thanks for the clear explanation..
Really h
View Tutorial By: Swj at 2011-12-30 05:55:26
5. this is no doubt a simple program for the learner
View Tutorial By: shweta at 2008-07-12 01:18:59
6. Easy to understand.......
View Tutorial By: Priya at 2013-02-15 10:55:52
7. Hiiii Al, I am Ashraf Abu Aisheh, i am working to
View Tutorial By: Ashraf Abu Aisheh at 2008-11-09 14:56:39
8. why java from two to the power 7 bit was upgraded
View Tutorial By: avinash at 2013-05-28 08:34:24
9. Mail me Codings in java of different programs
View Tutorial By: vikram at 2008-09-12 06:04:42
10. can u pls tell me how to retrive data from databas
View Tutorial By: yogesg desai at 2010-05-27 08:26:24 | http://java-samples.com/showcomment.php?commentid=34125 | CC-MAIN-2018-22 | refinedweb | 315 | 66.13 |
Previously in our Logger Refactor adventure, we extracted a class for logging to
System.out.
As mentioned; the focus of this part will be performing a similar operation for logging to
android.util.Log. (Note: This focus totally didn't happen...)
We can revisit the current state of
Logger.java as a reminder of what it currently looks like.
The big take away I see with this is that there's just the one logger. If we're planning on pulling new logger functionality into this... heh... Small steps. We gotta make SMALL steps. It's a hard bit to REALLY do.
What I was planning on doing was to encapsulate all the functionality of Logger into an SystemLogger inner class; then create a AndroidLogger inner class containing all that functionality; then pull out all common into the Logger class. That's what I ended up with when I did this before. I'll probably end up with it again. ...
As I'm sitting here pondering for a few moments. I'm looking at the logger and I see
final int level, final int loglevel. I have 2
ints; both representing a logging level. Then below I switch on on that level to return a value in
getLevelTag. Since
switches are bad; let's clean that guy up before we expect to drag more code into this.
I'm going to create a new class
LogLevel. Seems fitting given the fairly repetitive naming of the parameters.
public class LogLevel {}
Since our use of the log level is an
int and we
switch on it to get a
String, let's create a constructor with those.
public LogLevel(final int level, final String tag){ this.level = level; this.tag = tag; }
Much like the
Logger SystemOut; this is a stateless item that will be used repeatedly; a singleton makes sense.
The declaration for one of these looks like
public final static LogLevel VERBOSE = new LogLevel(Log.VERBOSE, "V");
and this will be done for each log level. These will get shown later.
Run Tests
The new classes aren't used; so it's really good that the tests still run.
To make use of the new
LogLevel objects; we need to update
FyzLog's log methods to pass the appropriate
LogLevel into the
println method.
Once we make that change it looks like
Not very exciting...
Run Tests .... DOESN'T COMPILE...
Oh, right... Then update the Logger to take this new parameter. The IDE tried to help me there.
We'll update the log level of the logging call. This makes it so we can no longer use
getLevelTag to get the tag; we'll need to expose that functionality on the
LogLevel.
public String tag(){ return tag; }
Now that we've updated
FyzLog, we have made changes to the code; the tests should be run.
Run Tests
Yay!
Now we can migrate the other logging level param for
println. As we make that change in
FyzLog we cause a compile error (between the '**')
private static void log(final int level, final String msgFormat, final Object... args) { if (**level >= logLevel** && msgFormat != null) {
We need to enable
LogLevel to compare. We want to create a method to let us know if the
level can log at the
logLevel.
if (level.logAt(logLevel) && msgFormat != null) {
For the short term; we'll turn the switch in the
log method into an
if-else so we can get be compilable again.
This also requires the android log branch to pass in the
LogLevel. I was hoping to hold off on that; but it's an Obvious Implementation (ish) and all sins are acceptable to get to green.
Whoa - We're having to update the test code with the change to the
FyzLog#logLevel variable.
Which is actually going to cause quite a bit of updating. My tests are refactored into a streamlined form that relies on the
int nature of the log levels.
We'll hold off updating the tests for the next post. For this; we didn't get our objective of extracting a
AndroidLog logger; but we created the
LogLevel class to encapsulate the logging levels and associated tag.
Our tests are broken since we've changed how the logging level is represented. If it was encapsulated earlier; we wouldn't have been able to write the tests that rely on the log level being an int.
The code compiles, the tests need love. We'll refactor the tests next week.
Part 1 - Part 2 - [Part 3] - Part 4 - Part 5 | https://quinngil.com/2017/01/18/android-logger-refactor-part-3/ | CC-MAIN-2021-10 | refinedweb | 757 | 74.69 |
Originally posted by Layne Lund: It sounds like there are problems when trying to read the numeric value for salary. Everything else is Strings as far as I can tell, sot here shouldn't be any major issues there. To find out the problem, you can call the ioException() method on the Scanner and print out a diagnostic:
// ...(snip)
System.out.print ( "Enter monthly salary: ");
double monthlySalary = input.nextDouble ();
System.out.println();
IOException ex = input.ioException();
ex.printStackTrace();
// ...(snip)
Try this and see if it gives any useful information. If you have any difficulty understanding the output, copy-and-paste it here and we'll help you decipher it. HTH Layne[/QB]
Originally posted by marc weber: My understanding is that Scanner is intended to read tokens, and by default these tokens are delimited by whitespace. My impression (from the API) is that the nextLine method is actually a way to skip past any remaining tokens on the current line, and reset to the next line. It then returns a String of everything that was skipped. In contrast, the nextDouble method does not reset to the next line. So after calling nextDouble, when you then call nextLine, it will skip over what comes after the double (looking for another token) and assign that String to firstName. Note that the code below prints the input. When it asks you for the salary, see what happens when you input "1234.56 fred". import java.util.Scanner;
public class Test {
public static void main(String[] args) {
Scanner input = new Scanner(System.in);
System.out.print("Enter salary: ");
double salary = input.nextDouble();
System.out.println("Input salary is: " + salary);
System.out.print("Enter first name: ");
String firstName = input.nextLine();
System.out.println("Input first name is: " + firstName);
System.out.print("Enter last name: ");
String lastName = input.nextLine();
System.out.println("Input last name is: " + lastName);
}
} So to get the code to work as you expect... After the nextDouble method is called, I think you need to call nextLine in order to reset to the next line. (Note that the return doesn't have to be assigned to anything. Simply adding a line of "input.nextLine();" should do it.) [ February 12, 2005: Message edited by: marc weber ] | http://www.coderanch.com/t/398529/java/java/keyboard-input-Scanner-object | CC-MAIN-2015-06 | refinedweb | 372 | 67.04 |
Welcome to Part 3 of Data Analysis with Pandas and Python. In this tutorial, we will begin discussing IO, or input/output, with Pandas, and begin with a realistic use-case. To get ample practice, a very useful website is Quandl. Quandl contains a plethora of free and paid data sources. What makes this location great is that the data is generally normalized, it's all in one place, and extracting the data is the same method. If you are using Python, and you access the Quandl data via their simple module, then the data is automatically returned to a dataframe. For the purposes of this tutorial, we're going to just manually download a CSV file instead, for learning purposes, since not every data source you find is going to have a nice and neat module for extracting the datasets.
Let's say we're interested in maybe purchasing or selling a home in Austin, Texas. The zipcode there is 77006. We could go to the local housing listings and see what the current prices are, but this doesn't really give us any real historical information, so let's just try to get some data on this. Let's query for "home value index 77006." Sure enough, we can see an index here. There's top, middle, lower tier, three bedroom, and so on. Let's say, sure, we got a a three bedroom house. Let's check that out. Turns out Quandl already provides graphs, but let's grab the dataset anyway, make our own graph, and maybe do some other analysis. Go to download, and choose CSV. Pandas is capable of IO with csv, excel data, hdf, sql, json, msgpack, html, gbq, stata, clipboard, and pickle data, and the list continues to grow. Check out the IO Tools documentation for the current list. Take that CSV and move it into the local directory (the directory that you are currently working in / where this .py script is).
Starting with this code, loading in a CSV to a dataframe can be as simple as:
import pandas as pd df = pd.read_csv('ZILL-Z77006_3B.csv') print(df.head())
Output:
Date Value 0 2015-06-30 502300 1 2015-05-31 501500 2 2015-04-30 500100 3 2015-03-31 495800 4 2015-02-28 492700
Notice that we have no decent index again. We can fix that like we did before doing:
df.set_index('Date', inplace = True)
Now, let's say we want to send this back to a CSV, we can do:
df.to_csv('newcsv2.csv')
We only have the one column right now, but if you had many columns, and just wanted to send one, you could do:
df['Value'].to_csv('newcsv2.csv')
Remember how we graphed multiple, but not all, columns? See if you can guess how to save multiple, but not all, columns.
Now, let's read that new CSV in:
df = pd.read_csv('newcsv2.csv') print(df.head())
Output:
Date Value 0 2015-06-30 502300 1 2015-05-31 501500 2 2015-04-30 500100 3 2015-03-31 495800 4 2015-02-28 492700
Darn, our index is gone again! This is because CSV has no "index" attribute like our dataframe does. What we can do, is set the index on import, rather than importing and then setting the index. Soemthing like:
df = pd.read_csv('newcsv2.csv', index_col=0) print(df.head())
Output:
Value Date 2015-06-30 502300 2015-05-31 501500 2015-04-30 500100 2015-03-31 495800 2015-02-28 492700
Now, I do not know about you, but the name "value" is fairly worthless. Can we change this? Sure, there are many ways to change the column names, one way is:
df.columns = ['House_Prices'] print(df.head())
Output:
House_Prices Date 2015-06-30 502300 2015-05-31 501500 2015-04-30 500100 2015-03-31 495800 2015-02-28 492700
Next, we can try to save to csv like so:
df.to_csv('newcsv3.csv')
If you look at the CSV there, you should see it has the headers. What if you don't want headers? No problem!
df.to_csv('newcsv4.csv', header=False)
What if the file doesn't have headers? No problem
df = pd.read_csv('newcsv4.csv', names = ['Date','House_Price'], index_col=0) print(df.head())
Output:
House_Price Date 2015-06-30 502300 2015-05-31 501500 2015-04-30 500100 2015-03-31 495800 2015-02-28 492700
These were the basics of IO an some options you have when it comes to input and output.
One interesting thing is the use of Pandas for conversion. So, maybe you are inputting data from a CSV, but you'd really like to display that data to HTML on your website. Since HTML is one of the datatypes, we can just export to HTML, like so:
df.to_html('example.html')
Now we have an HTML file. Open it up, and boom you have a table in HTML.
Note, this table is automatically assigned the class of "dataframe." This means you can have custom CSS to handle for dataframe-specific tables!
I particularly like to use Pandas when I have an SQL dump of data. I tend to pour the database data right into a Pandas dataframe, perform the operations that I want to perform, then I display the data in a graph maybe, or I otherwise serve the data in some way.
Finally, what if we want to actually rename just one of the columns? Earlier, you were shown how to name all columns, but maybe you just want to change one without having to type all the others out. Easy enough:
df = pd.read_csv('newcsv4.csv', names = ['Date','House_Price']) print(df.head()) df.rename(columns={'House_Price':'Prices'}, inplace=True) print(df.head())
Output:
Date House_Price 0 2015-06-30 502300 1 2015-05-31 501500 2 2015-04-30 500100 3 2015-03-31 495800 4 2015-02-28 492700 Date Prices 0 2015-06-30 502300 1 2015-05-31 501500 2 2015-04-30 500100 3 2015-03-31 495800 4 2015-02-28 492700
So here, we first imported the headless file, giving the column names of Date and House_Price. Then, we decided, nope we want to call House_Price just Price instead. So, we used df.rename, specifying that we wanted to rename columns, then, in dictionary form, the Key is the original name and the value is the new name. We finally use inplace=True so the original object is modified. | https://pythonprogramming.net/input-output-data-analysis-python-pandas-tutorial/?completed=/basics-data-analysis-python-pandas-tutorial/ | CC-MAIN-2021-39 | refinedweb | 1,089 | 81.93 |
Asynchronous Web Services Invocation in .NET Framework 2.0
Synchronous communication between .NET applications and Web services makes the user wait while each Web service processes requests and returns results. This can have a severe impact on the performance of the .NET application. Typically, a distributed .NET application requires information from multiple Web services. If the application performs the entire process of invoking the Web services synchronously, a client must wait not only until every Web service provider is contacted, but also through any connection or service delays.
Asynchronous Web services invocation solves this performance issue and enhances the end user experience by increasing server efficiency. With the introduction of .NET Framework 2.0, Microsoft has greatly enhanced the support for asynchronous Web services invocation by introducing a new event-based programming model. This article examines this new feature and demonstrates how to take advantage of it to create feature-rich and effective .NET applications. It also shows how to perform data binding directly with the results the Web service returns.
Event-Based Asynchronous Programming
Previous versions of the .NET Framework used the BeginInvoke/EndInvoke methods to invoke a Web service asynchronously. Version 2.0 adds a new way to asynchronously invoke Web services: an event-based asynchronous programming model. It enables this new event programming model through the creation of properties and methods on the client proxy class.
To follow the demonstration for implementing this model, you must first create a simple Web service that then can be invoked asynchronously from the client application. The following section walks you through the process.
Create a Simple Web Service
Open Visual Studio 2005 and select New->Web Site from the File menu. In the New Web Site dialog box, select ASP.NET Web service from the list of project templates, and specify the name of the Web service as AsyncWebServices.
Once you have created the Web site, add a new Web service file named HelloService.asmx to the project. If you open the file and view the code, you will find a method named HelloWorld already placed in the code-behind file of the Web service. For the purposes of this example, just utilize this default sample method. Here is the code from the code-behind file of the HelloService:
using System.Web; using System.Collections; using System.Web.Services; using System.Web.Services.Protocols; [WebService(Namespace = "")] [WebServiceBinding(ConformsTo = WsiProfiles.BasicProfile1_1, EmitConformanceClaims = true)] public class HelloService : System.Web.Services.WebService { [WebMethod] public string HelloWorld() { return "Hello World"; } }
Note: When you create a Web service through Visual Studio 2005, the code-behind file for the Web service is automatically placed in the App_Code directory, which is a new directory in ASP.NET 2.0 that acts as a container for all the reusable classes in an ASP.NET Web application.
Now that you've created the Web service, if you right click on it and select View in Browser, you will see the screen shown in Figure 1.
Figure 1. View Newly Created Web Service in Browser
If you click on the HelloWorld method, you will see the screen shown in Figure 2.
Figure 2. View HelloWorld Method in Browser
Clicking on the Invoke button in the Figure 2 results in the screen shown in Figure 3.
Figure 3. Invocation of HelloWorld Method
Figure 3 shows the output produced by the HelloWorld method. Now that you have implemented and tested the Web service, you can move on to the client application that will consume the Web service.
Page 1 of 3
| http://www.developer.com/net/net/article.php/3481691/Asynchronous-Web-Services-Invocation-in-NET-Framework-20.htm | CC-MAIN-2014-10 | refinedweb | 589 | 57.98 |
This plugin adds a local database to ONIS 2.6 that is used to
import DICOM studies, series and images on your local hard disk. You can then
browse this local database to find, open, transfer or export images. The
database can also store your annotations and reports.
This database is based on MS-ACCESS. It is one of the
simpliest and most flexible database solutions on the market today. It has a 2GB
maximum size limitation, which is sufficient for storing tens of thousand of
DICOM images.
This plugin will add a "Local" source in ONIS 2.6 and a "Local
server" page in the preferences panel. As soon as the plugin is loaded, you can
start to import, transfer, export or open images from the local source. This
database will also be the source for other plugins such as the DICOM Server or the
Remote Server plugins.
This plugin is free of charge and is included in all of the three packages (Free edition,
Professional, and Ultimate packages). If you need a store hundreds of thousands or millions of DICOM images (typically if you want to use ONIS 2.6 as your DICOM server and/or Remote server), you will probably want to use a more professional and reliable database such as MSSQL Server 2008. | http://www.onis-viewer.com/PluginInfo.aspx?id=46 | CC-MAIN-2018-09 | refinedweb | 217 | 64.61 |
{-#LANGUAGE MultiParamTypeClasses, FunctionalDependencies, FlexibleInstances, UndecidableInstances #-} {- | 'last', . -} module Data.FixedList ( -- * Types and Classes Cons(..) , Nil(..) , FixedList , Append (..) -- * Baisc Functions that are not found in 'Traversable' or 'Foldable' , reverse , length , last , init , unit , subLists , fromFoldable , fromFoldable' -- * Type synonyms for larger lists , FixedList0, FixedList1, FixedList2, FixedList3, FixedList4 , FixedList5, FixedList6, FixedList7, FixedList8, FixedList9 , FixedList10, FixedList11, FixedList12, FixedList13, FixedList14 , FixedList15, FixedList16, FixedList17, FixedList18, FixedList19 , FixedList20, FixedList21, FixedList22, FixedList23, FixedList24 , FixedList25, FixedList26, FixedList27, FixedList28, FixedList29 , FixedList30, FixedList31, FixedList32 ) where import Control.Applicative import Control.Monad hiding (sequence) --hiding for haddock import Data.Foldable import Data.Traversable import Data.Monoid import Data.Maybe import qualified Data.List import Prelude hiding (head, tail, foldr, sum, sequence, reverse, length, last, init) data FixedList f => Cons f a = (:.) { head :: a, tail :: (f a) } deriving (Eq, Ord) data Nil a = Nil deriving (Eq, Ord) infixr 5 :. instance (FixedList f, Show a) => Show (Cons f a) where show x = "|" ++ show (toList x) ++ "|" instance Show (Nil a) where show Nil = "|[]|" -- | Just a restrictive typeclass. It makes sure ':.' only takes FixedLists as it's second parameter -- and makes sure the use of fromFoldable's in reverse, and init is safe. class (Applicative f, Traversable f, Monad f) => FixedList f instance FixedList f => FixedList (Cons f) instance FixedList Nil -- The only very bad ugly, everything else is haskell98 class Append f g h | f g -> h, f h -> g where append :: f a -> g a -> h a instance (FixedList f, FixedList c, Append f b c) => Append (Cons f) b (Cons c) where append (x :. xs) ys = x :. (xs `append` ys) instance Append Nil a a where append Nil ys = ys reverse :: FixedList t => t a -> t a reverse xs = fromFoldable' $ Data.List.reverse $ toList xs length :: Foldable t => t a -> Int length xs = Data.List.length $ toList xs -- | Returns the last element of the list last :: Foldable t => t a -> a last xs = Data.List.last $ toList xs -- | Returns all but the last element of the list init :: FixedList f => Cons f a -> f a init xs = fromFoldable' $ Data.List.init $ toList xs -- | Constructs a FixedList containing a single element. -- Normally I would just use pure or return for this, -- but you'd have to specify a type signature in that case. unit :: a -> Cons Nil a unit a = a :. Nil -- | Given a list, returns a list of copies of that list but each with an element removed. -- for example: -- -- > subLists (1:. 2:. 3:. Nil) -- -- gives: -- -- > |[|[2,3]|,|[1,3]|,|[1,2]|]| -- subLists :: FixedList f => Cons f a -> Cons f (f a) subLists xs = fromFoldable' $ fmap fromFoldable' $ subLists' $ toList xs where subLists' a = zipWith (++) (Data.List.inits a) ( Data.List.tail $ Data.List.tails a) -- | Converts any Foldable to any Applicative Traversable. -- However, this will only do what you want if 'pure' gives you the -- shape of structure you are expecting. fromFoldable :: (Foldable f, Applicative g, Traversable g) => f a -> Maybe (g a) fromFoldable t = sequenceA $ snd $ mapAccumL f (toList t) (pure ()) where f [] _ = ([], Nothing) f (x:xs) _ = (xs, Just x) -- | This can crash if the foldable is smaller than the new structure. fromFoldable' :: (Foldable f, Applicative g, Traversable g) => f a -> g a fromFoldable' a = fromJust $ fromFoldable a instance FixedList f => Functor (Cons f) where fmap f (a :. b) = f a :. fmap f b instance FixedList f => Foldable (Cons f) where foldMap f (a :. b) = f a `mappend` foldMap f b instance FixedList f => Traversable (Cons f) where traverse f (a :. b) = (:.) <$> f a <*> traverse f b instance FixedList f => Monad (Cons f) where return x = x :. return x (a :. b) >>= k = head (k a) :. (b >>= (tail . k)) instance FixedList f => Applicative (Cons f) where pure = return (<*>) = ap instance Functor Nil where fmap _ Nil = Nil instance Foldable Nil where foldMap _ Nil = mempty instance Traversable Nil where traverse _ Nil = pure Nil instance Monad Nil where return _ = Nil Nil >>= _ = Nil -- should I define this as _ >>= _ = Nil ? would this -- have the right behavior with bottom? instance Applicative Nil where pure = return (<*>) = ap instance (Num a, FixedList f, Eq (f a), Show (f a)) => Num (Cons f a) where a + b = pure (+) <*> a <*> b a - b = pure (-) <*> a <*> b a * b = pure (*) <*> a <*> b negate a = pure negate <*> a abs a = pure abs <*> a signum = fmap signum fromInteger = pure . fromInteger instance (Fractional a, FixedList f, Eq (f a), Show (f a)) => Fractional (Cons f a) where a / b = pure (/) <*> a <*> b recip a = pure 1 / a fromRational = pure . fromRational | http://hackage.haskell.org/package/fixed-list-0.1.5/docs/src/Data-FixedList.html | CC-MAIN-2015-27 | refinedweb | 741 | 70.84 |
digitalmars.D - Re: Reddit: why aren't people using D?
- Kagamin <spam here.lot> Jul 28 2009
Andrei Alexandrescu Wrote:Might be because (a) we aren't getting our priorities right, (b) we ascribe more to properties than what the compiler really makes of them. At the end of the day, a property is a notational convenience. Instead of writing: obj.set_xyz(5); int a = obj.get_xyz(); properties allow us to write: obj.xyz = 5; int a = obj.xyz; Just to be 100% clear, I agree that the convenience is great. But I don't know why the hell I need to learn a whole different syntax for *defining* such things, when the compiler itself doesn't give a damn - it just blindly rewrites the latter into something like the former.
May be namespace states your complaints. It doesn't add a whole new syntax, only kinda one extra level of indirection, naturally reusing existing feature of operator overloading, and I think it should look natural to compiler too: it's the same as operator overloading, compiler already does it. And calling properties is only one aspect of the problem. The other is declaration. You heard it: implicit backing storage, default trivial implementation, type consistency. We don't have problems with calling properties. In fact complaints about current design originates from problems with property declaration: the problem is compiler doesn't know what is a function and what is a property.
Jul 28 2009 | http://www.digitalmars.com/d/archives/digitalmars/D/Re_Reddit_why_aren_t_people_using_D_94069.html | CC-MAIN-2014-10 | refinedweb | 244 | 58.58 |
.
In the original game you play Guybrush Threepwood, a would-be pirate who must learn to become a great swordfighter if he's ever to realize his pirate dreams. In this case, however, the key to great sword fighting is in the mind, not in the hands, because the game is all about insults and comebacks—the fighter with the sharper tongue wins.
My version of the game is a blend of Rock-Paper-Scissors meets Memory meets Fencing. To begin, you select your character and your (computer) opponent (see Figure 1). Next, you select an insult from a list of insults to throw your opponent's way. If your opponent has an appropriate comeback, he will take the lead and throw an insult your way. Then it's your turn to select an appropriate comeback to reclaim the lead. If you select the wrong comeback, you lose. However, if you can hold your ground, you'll advance to higher levels, and will be able to unlock hidden characters. The final character in the game is a boss, who uses a special set of insults. (However, the same comebacks you've already learned will work against him. It's up to you to figure out which comeback goes with what insult.)
Tip: When playing the game, losing a lot in advance is key. Using wrong comebacks over and over will enable you to learn even more. When you meet the boss, your insults are useless to you. It all depends on the comebacks. Slow and steady wins the race.
Figure 1. Select your character.
Note: In Figure 1, Insult Dueler is running on top of the browser.
One of the cool things about Adobe AIR is that you can choose whether you want your app to use the standard window chrome (which has the look and feel of any other application window on your Windows or Mac OS system) or custom chrome. I'm a custom man myself, so I chose to go the custom chrome route—which also meant that I was responsible for adding my own Close and Minimize buttons (the Maximize button is not necessary for this game).
Luckily, all of the necessary code is built into Adobe AIR. For example, if you want to add minimize functionality to a window, all you need to do is this:
someDisplayObject.stage.nativeWindow.minimize();
Another fun feature of custom chrome is that you can build an exotic-looking window with its own transparency (see Figure 1). For this game, I even added a slider to control the alpha of the backmost movie clip, so users can choose their own transparency level.
But wait! There's more!
It's one thing to manipulate an app window—to move it around or make it semi-transparent. It's another matter entirely to remember the location and appearance of the application when the user quit it. To pull this off, you'll need to write a file to the user's hard drive that stores the x and y coordinates as well as the alpha value of the window. For this game, I also saved the win/loss record—to add replay value. I saved all this information in an XML file, since it's pretty easy to read. My default XML file looks like this:
<prefs> <stagePosition x='0' y='0' alpha='1' /> <record wins='0' losses='0' /> </prefs>;
When the application first opens, it'll attempt to read the file like this:
var s:FileStream = new FileStream(); var f:File = File.applicationStorageDirectory; // this is a reference to a predetermined directory. check the air api for more. f = f.resolvePath(Constants.FILE_PREFS_XML); // basically converts a String to something the File class will understand. s.addEventListener(Event.COMPLETE,_onPrefsReadSuccess,false,0,false); s.addEventListener(IOErrorEvent.IO_ERROR,_onPrefsReadIOError,false,0,false); // opens, reads and when complete fires Event.COMPLETE (aka our _onPrefsLoaded function). s.openAsync(f,FileMode.READ);
That's it. Seven lines of code (minus comments) attempts to open a file. If it opens, you'll get
_onPrefsReadSuccess; otherwise you get
_onPrefsReadIOError. The IO error will realistically happen if the file is corrupt or if it doesn't exist at all—in either case, the game will simply begin to animate artwork. The success function is much more interesting:
private function _onPrefsReadSuccess($evt:Event):void { Out.status(this,"_onPrefsLoaded"); // hurrah! I know this is a returning user because the prefs xml has been created. let's set the user back where they left off. var xml:XML = new XML($evt.target.readUTF()); _mc.stage.nativeWindow.x = xml.stagePosition.@x; _mc.stage.nativeWindow.y = xml.stagePosition.@y; _wins = xml.record.@wins; _losses = xml.record.@losses; Out.info(this,"Stage Position: " + _mc.x.toString() + " x " + _mc.y.toString()); // prefs are set, now I can animate the screen in. _animateIn(); };
The app moves the stage window position to the number described in XML. It also stores the
_wins and
_losses as variables. Then it proceeds to display the content. What happens, after a game is won or lost, or the window gets dragged around? Have a look:
private function _updatePrefs():void { Out.status(this,"_updatePrefs"); // for updating, I'll do a synchronized file write. since this is a small file, it shouldn't require too much time to process. var s:FileStream = new FileStream(); var f:File = File.applicationStorageDirectory; f = f.resolvePath(Constants.FILE_PREFS_XML); // open the file here s.open(f,FileMode.WRITE); // create a copy of the schema I'll use for preferences xml, so I'm not editing the schema itself. var xml:XML = Constants.SAVED_PREFS.copy(); // populate the data I need. xml.stagePosition.@x = _mc.stage.nativeWindow.x; xml.stagePosition.@y = _mc.stage.nativeWindow.y; xml.record.@wins = _wins; xml.record.@losses = _losses; Out.debug(this,"Prefs written as follows:"); Out.debug(this,xml.toXMLString()); // and finally I write. s.writeUTF(xml); s.close(); };
This time set
FileStream to open the file as writable (the
s.open() line). Then do the opposite of the read: Grab the saved variables (window position, wins and losses) and write the XML data as UTF bytes. Easy as pie.
Now that I have a cool custom application, I shall fill that window with my game.
As mentioned earlier, this game is all about insults and comebacks. As you progress you unlock content. Although this is an extraordinarily simple game, the key to realizing this is keeping my code organized. I don’t want to wade through a hundred lines of game code to determine when a user selects a character. I want my character selection to be separate from my game. That way, if there’s a bug with a character selection screen I know just where to look. I want all the major pieces of functionality divided in a way that will allow me to describe specific functionality.
I follow a very loose organizational model for my code. The engine that drives this game is broken up to serve three types of classes: screens, game data, and user interface (UI).
To tie it all together, I have a Document Root class called
Main. This class instantiates (new
MyClass()) all of the other classes and listens for any events.
Main is the hub that tells the other classes how to work together. In addition,
Main is responsible for any preliminary loading.
The
Constants class contains all of the permanent, never-changing variables needed for the game to work. This class contains the entire script for the game in array format. The
INSULTS array looks like this:
public static const INSULTS :Array = [ then you.", "You're as repulsive as a monkey in a negligee.", "You're the ugliest creature I've ever seen in my life!", "I'll hound you night and day!"];
Every other class imports
Constants. If I want to output "You fight like a dairy farmer.", I can type
trace(Constants.INSULTS[7]).
I have a similar array called
COMEBACKS and yet another called
BOSS_INSULTS. I'm not going to give those away so easily, though—you'll have to play the game to find the comebacks.
To offer the user a clue as to when the computer answers with an incorrect comeback, I added an array of guaranteed wrong comebacks. These comebacks are always wrong. Use 'em and you'll lose every time.
public static const GENERIC_WRONG_COMEBACKS :Array = [ "I am rubber you are glue.", "Oh yeah?", "This is going to hurt, isn't it?", "Uh... chicken parm?", "Golly, that's a good one!", "Look behind you! A three-headed monkey!", // these next two are really, really cheesy. I'm sorry. I can't resist the urge to cheese it up. "I'm terrible at this. I'm going to quit swordfighting and learn ActionScript instead.", "Well, I'm going to lose this battle... but at least I'll lose it in a transparent window." ];
Constants also contains an array called
CHARACTER_DATA. CHARACTER_DATA is the single most important variable in the site. Each value in the array is an object with the following properties:
Constants.CHARACTER_DATA[0].KEYshould be 0. This way I have the flexibility to reference an object by its place in the array or by the object. I know what each character object's key is because I know that I'm never going to move the array around. It will remain constant. Hence the aptly named class!
Mainloads this file and passes it to the game screen.
Mainloads this and passes it to the menu screen.
In code, it looks a little like this:
public static var KEY_BABAR:int = 0; // now I can reference Guybrush from Constants by using Constants.KEY_BABAR public static var CHARACTER_DATA :Array = [{ KEY: KEY_BABAR, SWF: "babar.swf", PORTRAIT: "babar.jpg", SHORT_NAME: "Babar", LONG_NAME: "Babar, Lord of the French Bulldogs”, TEXT_FOREGROUND: 0x000000, TEXT_BACKGROUND: 0xFFFFFF, BOSS: false, UNLOCKED: true }, ];
All I need to do is add additional objects for all the other characters.
You might be wondering about
Constants.KEY_BABAR. I know I can access all the Guybrush data from
Constants.CHARACTER_DATA[0], so why bother with the extra variable? There are two reasons. First, to protect myself. If I accidentally write
Constants.CHARACTER_DATA[1] somewhere, the site uses the wrong character data, creating a bug. However, if I accidentally type
Constants.CHARACTER_DATA[Constants.KEY_BABARR] (with an extra R), that variable doesn't exist and the compiler will tell me there's an error. Way easier to diagnose.
The second reason is to make the code easier to read for other developers. Don't you hate looking at code like this:
var a1 = b26; b26 += mClip352["stuff_" + temp];
To another programmer, this is a meaningless mess of letters and numbers.
KEY_BABAR, although more to type, is more eloquent. Another programmer will see that and know right away that whatever is going on is happening specifically to Guybrush.
The insults, comebacks, and characters are all stored in arrays. Because of this structure, I know that all the scoring and game progress is entirely based around these arrays. Thus I'll know when a new insult or comeback appears, all I need to do is store the key somewhere for retrieval the next time the user opens the game. But how can I manage saving progress?
The Adobe AIR framework comes complete with SQLite support. Basically, this means I have a collection of classes that allows me to create and control a local database entirely from ActionScript. I use this database to do the following:
The only issue is that this whole SQLLite-create-and-control-my-own-database is all brand new. Which means… uh oh… new syntax. Scary! Deep breaths…
_sql = new SQLConnection(); _sql.addEventListener(SQLEvent.OPEN,_onSQLConnectSuccess,false,0,true); _file = File.applicationStorageDirectory; _file = _file.resolvePath(Constants.FILE_DATABASE); _sql.openAsync(_file, "update");
… wait a minute. That's it?
Yep. Well, not quite. But that's the basics. In five lines of code, I say the following:
_onSQLConnectSuccess.
Constants.FILE_DATABASEis set to
insultDueler.db.
Okay… so, what if the database file doesn't exist yet? How does Adobe AIR know to create it? The syntax above will automatically create the file if it doesn't exist. If you didn't want that, modify the code as follows (changes in bold):
_sql = new SQLConnection(); _sql.addEventListener(SQLEvent.OPEN,_onSQLConnectSuccess,false,0,true); _sql.addEventListener(SQLErrorEvent.ERROR,_onSQLConnectFail,false,0,true); _file = File.applicationStorageDirectory; _file = _file.resolvePath(Constants.FILE_DATABASE); _sql.open(_file,false);
Now the app will fire a SQLErrorEvent if it can't find the database file. And the
_open method is passed a false, specifically telling it not to auto-create the database file.
_onSQLConnectFail looks like this:
private function _onSQLConnectFail($evt:SQLErrorEvent):void { Out.debug(this,"_onSQLConnectFail"); _isFirstInit = true; _sql.openAsync(_file, "create"); };
I set an
_isFirstInit Boolean to true, which means I can open the database again (this time auto-creating it). I know it's the first time I'm opening it, so I'll set it to be the smallest database possible (512 is the smallest alloId) and set it so it doesn't automatically clean excess data (that's the false). Since I'm already listening for
_onSQLConnectSuccess, that'll be my next stop.
private function _onSQLConnectSuccess($evt:SQLEvent):void { var statement:SQLStatement = new SQLStatement(); statement.sqlConnection = _sql; if(_isFirstInit) statement.text = "CREATE TABLE " + Constants.TABLE_INSULTS + "(id INTEGER PRIMARY KEY AUTOINCREMENT,key INTEGER)"; else statement.text = "SELECT * FROM " + Constants.TABLE_INSULTS; Out.debug(this,"_onSQLConnectSuccess: " + statement.text); statement.addEventListener(SQLEvent.RESULT,_onSQLInsults,false,0,true); statement.execute(); };
So I've connected. Now I'll create a new SQL statement. I'll use this statement to tell the database what I need to do. If you've seen mySQL queries before, this should all look eerily familiar. If not, here's what I'm telling the database to do: If I just created the database, create a new table. This table will be named INSULTS (the variable I have in
Constants). The table will keep track of two variables: An ID that automatically counts up as I add more to the database (this is a pretty standard practice) and a KEY, which will be a variable I pass. Otherwise, I'll grab all of the rows in the existing table ("SELECT * FROM…") and wait for the result. When I get that data, I just need to turn it into an Array:
_knownInsults = []; var results:SQLResult = $evt.target.getResult(); if(results) for(var i:int=0;i<results.data.length;i++) { _knownInsults.push(results.data[i].key); }
When a new insult needs to be added to the database, I'll run this function:
public function addToTable($table:String,$key:Number):void { var statement:SQLStatement = new SQLStatement(); statement.sqlConnection = _sql; statement.text = "INSERT INTO " + $table + "(KEY) VALUES(" + $key + ")"; Out.debug(this,"addToTable: " + statement.text); statement.execute(); }; addToTable(Constants.TABLE_INSULTS,5); // this example would add a new row into my database. Assuming this is the first row ever, the ID will automatically be 1 and the KEY will be 5.
The next time the app opens, these same functions will run, populating my
_knownInsults array—all before the user sees a thing!
The menu screen is the screen containing all of the different selectable characters. I tried to emulate the interface of games like Street Fighter or Mortal Kombat. When in the menu screen is loaded, the user will see a list of portraits. Any of the characters still locked will have their portraits replaced by a generic lock icon and cannot be selected.
The movie clip passed to it contains two frame labels: IN and OUT (see Figure 2). These are called to play when the screen is meant to animate in and out respectively. At the end of each of these timelines is an animation event, a simple custom event I use to represent event hooks; a milestone on the Timeline.
Figure 2. The Timeline for the movie clip passed to the MenuScreen class.
In Figure 2, the ActionScript right before the OUT label looks as follows:
stop(); dispatchEvent(new AnimationEvent(AnimationEvent.ANIMATE_IN));
In my class, I map the event being dispatched (
AnimationEvent.ANIMATE_IN) to a method. This code is very simple:
public function animateIn():void { _mc.gotoAndPlay("IN"); _mc.addEventListener(AnimationEvent.ANIMATE_IN_START,_onAnimateInStart, false,0,true); _mc.addEventListener(AnimationEvent.ANIMATE_IN,_onAnimateIn,false,0,true); };
When the Timeline reaches Frame 23 (the frame that dispatches
AnimationEvent.ANIMATE_IN),
_onAnimateIn will execute and I'll know the movie clip is ready to fly.
For example, each of those movie clips you see in Figure 2 is a menuClip. When
_onAnimateIn is fired, I set all of these movie clips to
enabled = true, allowing the user to click on them.
In addition, these movie clips contain ROLLOVER, ROLLOUT labels, just like the IN and OUT labels, except this time, I'm going to use them for mouse events. Hence ROLLOVER and ROLLOUT. As they come in, they'll fire their own event (
MenuEvent), which
MenuScreen is also listening to. It reacts like this:
private function _onMenuItemPortrait($evt:MenuEvent):void { Out.info(this,"_onMenuItemPortrait"); // if the character is unlocked, display their portrait. otherwise, grab the "Locked" icon from the Library and display that. $evt.menuClip.empty.addChild(Constants.CHARACTER_DATA[$evt.key].UNLOCKED ? _portraitsArray[$evt.key].content : new Bitmap(new LibraryItem_Lock(1,1))); }; private function _onMenuItemAnimateIn($evt:MenuEvent):void { _menuDictionary[$evt.menuClip] = Constants.CHARACTER_DATA[$evt.key]; Out.info(this,"_onMenuItemAnimateIn"); // again, if the character is unlocked their button is active. otherwise set their button disabled. if(Constants.CHARACTER_DATA[$evt.key].UNLOCKED) { $evt.menuClip.btn.addEventListener(MouseEvent.CLICK,_onMenuItemClick,false,0,true); $evt.menuClip.btn.addEventListener(MouseEvent.ROLL_OUT,_onMenuItemRollOut,false,0,true); $evt.menuClip.btn.addEventListener(MouseEvent.ROLL_OVER,_onMenuItemRollOver,false,0,true); } $evt.menuClip.btn.enabled = Constants.CHARACTER_DATA[$evt.key].UNLOCKED; };
This translates as follows: When a
menuClip is ready to display a portrait (the JPG loaded in from
Main), attach it to a movie clip named empty. Then, when
menuClip is fully displayed, if that menu clip's corresponding character data (found in
Constants) is unlocked, listen for the mouse events so I know when the user rolls over, rolls out, and clicks. Otherwise, disable the button so the user can't possibly be confused into selecting a locked character.
When the user has selected a character,
MenuScreen tells
Main which user character and opponent character the game screen needs to display. This chain of events unfolds like this:
// this is in MenuScreen. Main is listening for ScreenEvent.GAME_START. dispatchEvent(new ScreenEvent(ScreenEvent.GAME_START,userKey,opponent.KEY)); // this is in Main. This fires when MenuScreen dispatches the ScreenEvent above. _curScreen is the GameScreen. private function _onGameStart($evt:ScreenEvent):void { _curScreen = Constants.SCREEN_GAME; _screens[_curScreen].onCharactersSelected($evt); };
Like.
Artificial intelligence for games is never about creating a realistic replica of what a user would do. It's about creating the illusion of a realistic replica of what a user would do. As simply as possible. In this case, the computer needs to be smart enough to occasionally spit the correct comeback at the user. But it also needs to be dumb enough to mess up. Otherwise the user would never be able to win, which doesn't sound like a very fun game to me.
For insults, this is easy enough—pick a random insult, remember its index (just like
_lastUserChoiceSelected) and display the text on the screen. For comebacks it's only slightly trickier:
var shouldOpponentBeCorrect:Boolean = _(_random(0,2) == 0); if(shouldOpponentBeCorrect) { _lastOpponentChoiceSelected = _lastUserChoiceSelected; _mc.opponentTxt_mc.glow_mc.txt.text = Constants.COMEBACKS[_lastOpponentChoiceSelected]; } else { _lastOpponentChoiceSelected = -1; _mc.opponentTxt_mc.glow_mc.txt.text = Constants.GENERIC_WRONG_COMEBACKS[_random(0, Constants.GENERIC_WRONG_COMEBACKS.length-1)]; } var t:Timer = new Timer(_mc.opponentTxt_mc.glow_mc.txt.text.length * Constants.READ_SPEED); t.addEventListener(TimerEvent.TIMER,_onOpponentTimer,false,0,true); t.start();
_random() is a utility function I created to pick a random number between a minimum and maximum range. In this case, the minimum is 0, the maximum is 2.
shouldOpponentBeCorrect is a Boolean set to true only if that random number is 0. In other words, roughly 33% of the time the computer will answer with the correct insult. Then I display the text on the screen for the opponent the exact same way I did for the user. That's all. Artificial Intelligence indeed.
Suppose the gamer has chosen "You fight like a dairy farmer." I know that translates to Insult #7 in my array, counting from 0. I have this saved as
_lastUserChoiceSelected. I know the computer has just selected a comeback, the index of which I have stored as
_lastOpponentChoiceSelected.
Scoring is as simple as this:
// award 1 point to whoever won the round if(_lastOpponentChoiceSelected == _lastUserChoiceSelected) { if(!_knownComebacksObject["key_" + _lastOpponentChoiceSelected.toString()] && !Constants.CHARACTER_DATA[_opponentKey].BOSS) { _knownComebacks.push(_lastOpponentChoiceSelected); // now I can use this comeback! _knownComebacksObject["key_" + _lastOpponentChoiceSelected.toString()] = true; dispatchEvent(new DatabaseSaveEvent(DatabaseSaveEvent.SAVE,_lastOpponentChoiceSelected, Constants.TABLE_COMEBACKS)); // that's it. Main will ask the Database to save the comeback. } _score--; } else _score++; _userMC.gotoAndPlay("PARRY"); _opponentMC.gotoAndPlay("PARRY");
Add to
_score if the user won the round, subtract if the opponent did. In addition, if the user lost the round but didn't know the comeback (which I can check from
_knownComebackObjects), add that comeback to the
_knownComebacks array. Now the user can use that comeback.
The next round is the same flow reversed:
_lastOpponentChoiceSelectedis saved.
LibraryItem_GameOptions, this time with the text of all known comebacks.
Detecting Adobe AIR is essential for several reasons. First, a Flash app assumes it's been built for the browser. If you try to use an Adobe AIR method when not in Adobe AIR, your application will break. At Big Spaceship I built a simple utility class which I distribute called
Environment. In it, I have a static function that returns a Boolean: true if this is in Adobe AIR, false if not:
public static function get IS_IN_AIR():Boolean { return Capabilities.playerType == "Desktop"; };
IS_IN_AIR enables you to detect this pitfall at runtime, so your Adobe AIR methods only happen when needed.
In addition sometimes the flow of testing in the IDE is different than in final deployment. For example, in the IDE you can't create an XML file if it doesn't already exist. If you can detect that you're not in AIR, then you can program a way to bypass loading an XML file you know won't be there.
So my game works. Now I need to get it out the world. How do I get it out to the world? (Write a tutorial for Adobe, that’s how.)
I built my game with the SWF file set to export for Flash Player 9. With the Adobe AIR extension in Flash CS3, I can change this setting to Adobe AIR right here. But I’ve got a nifty idea.
The AIR extension in Flash CS3 will export an AIR file directly. But then in order to test the file, I have to install the application and see if it’s running correctly. I’m a speed demon. I want to test in the IDE first, and then package for Adobe AIR. So what I did was set my main.fla—the the file that contains all the game mechanics and artwork—to export a normal, regular SWF. Now I can see the game immediately. My
Environment.IS_IN_AIR Boolean is coming up false (since I’m not in Adobe AIR), but that’s okay. I can see the game work and diagnose bugs quickly.
Note: Why did I bother with a separate FLA for Adobe AIR, when I could just set main.fla to export as such? proxy.fla is my know-all, end-all AIR FLA file, that's why. It loads in main.swf, but doesn't actually care what main.swf does. I can reuse it over and over for porting SWF files to Adobe AIR. This gives me the flexibility to build my application for the browser and the desktop simultaneously!
Environment.IS_IN_AIR will be false in the browser, so my app won't even bother to try to write to the file system. Talk about standards, right?
So okay, I’ve gotten the game pretty tight. Now it’s time to test in AIR. I created
proxy.fla for just this moment.
proxy.fla is the exports as Adobe AIR, and the only thing it does is load in
main.swf. The
proxy.fla file’s Document class reads like this:
package com.kosoy.insultdueler { // native AS3 import flash.net.URLRequest; import flash.display.Loader; import flash.display.MovieClip; // AIRProxy is the Document Root class of proxy.fla. public class AIRProxy extends MovieClip { public function AIRProxy() { var l:Loader = new Loader(); addChild(l); l.load(new URLRequest("main.swf")); }; }; }
proxy.swf loads in
main.swf. Thus
main.swf is in Adobe AIR. Even better, I could actually reuse
proxy.fla for any SWF file I want to convert to Adobe AIR for the rest of my life. It’s super modular. I just need to name the SWF file
main.swf.
It is extraordinarily difficult to test in Adobe AIR. It’s horrible. Painful. Don’t even try it.
I test the movie as I normally would. Instantly things are wildly different. My application lives! It’s got the pretty custom chrome, the transparency, the Minimize and Close buttons work, it even remembers my win-loss record and alpha and…
Wait a minute. This isn’t a real application yet. I haven’t packaged it and installed it. Where is it saving the database and the XML file to? Wherever
proxy.fla is located, that’s where. This is great! Now I can open the XML file and see what is actually saved! I can see all the files coming to life. I didn’t have to do a single thing differently!
Okay, so I’ve tested in the IDE and I’ve tested in Adobe AIR. Now at long last I’m ready to build the final AIR file. Like I mentioned, the proxy FLA file is set to export to Adobe AIR. By selecting Commands > AIR – Application and Package Settings I can set everything I need to create an AIR file. In this case, I added the media and swf directories, set some custom icons and add a little bit of copy information. The end result looks like Figure 5.
Figure 5. Insult Dueler Application and Package Settings
ActionScript 3.0 is a big step forward for Flash developers. For the first time, we have a "mature" programming language capable of reading any kind of data we can imagine. Adobe AIR takes things a step further, enabling writing all of that data to the desktop. For the first time, Flash designers and developers can start to take their expertise to the desktop.
Adobe AIR empowers developers to build applications of incredible complexity with ease. In one fell swoop I'm reading and writing an XML file, adding and reading data from a SQL database and building a cross-platform game. And all I had to learn was a tinsy bit more ActionScript 3.0. I even built in my own custom icons!
Oh, and in case anybody is wondering, as of this writing my win-loss record is 12 to 20. You would think I'd be better at this game, seeing how I'm the one that built it. Ah well. Only one way to get better, right? Just keep playing and playing and playing and playing…
If you want to learn more about ActionScipt 3.0, Grant Skinner and Senocular have blog-tomes filled with wonderful knowledge. The ActionScript 3 Language Reference is an obvious necessity for wading through the new syntax. And of course, Big Spaceship Labs is a wonderful resource from some of the finest developers on the planet.
For more inspiration, check out the sample apps in the Adobe AIR Developer Center for Flash, Flex, and HTML/Ajax.
To get started building Flash apps on Adobe AIR go to the Getting Started section of the Adobe AIR Developer Center for Flash or dig into Developing Adobe AIR Applications with Adobe Flash CS3.. | http://www.adobe.com/devnet/air/flash/articles/insult_dueler_print.html | crawl-002 | refinedweb | 4,689 | 59.19 |
The days when an application was a simple process that ran entirely locally are long gone. Most systems these days are complex beasts that reach out to a variety of services which can be hosted both locally and remotely. As the number of links in the chain grows, so does the potential for a performance problem to surface. And as we delegate more and more work to 3rd party services instead of owning them, sometimes the best we can hope for is to keep a very close eye on how they’re behaving.
The internal components we also rely on will continually get upgraded with new features and bug fixes so we need to make sure that any change in their performance characteristics are still within our expected tolerances. Even during day-to-day operation there will be spikes in volumes and anomalous behaviour in both our own code and those of our upstream and downstream dependencies that will need to be investigated.
If you’ve ever had a support query from a user that just says “the system is running like a dog” and you have no performance data captured from within the bowels of your application then you’ve got an uphill battle. This article explores what it takes to add some basic instrumentation to your code so that you can get a good idea of what is happening in and around your system at any time.
System-scale monitoring
Your operations team will no doubt already be monitoring the infrastructure, so why isn’t that enough? Well, for starters they are almost certainly only monitoring your production real estate – your development and test environments are likely to be of little importance to them. And yet those are the environments where you want to start seeing any performance impact from your changes so that they can be dealt with swiftly before reaching production.
The second reason is that they only have a 1,000 foot view of the system. They can tell you about the peaks and troughs of any server, but when an anomaly occurs they can’t tell you about which service calls might have invoked that behaviour. Databases in particular have a habit of behaving weirdly as the load spikes and so knowing which query is being blocked can give you clues about what’s causing the spike.
What to measure
Sir Tony Hoare tells us that “premature optimisation is the root of all evil” [Hoare] and therefore the first rule of thumb is to measure everything we can afford to. In a modern system where there are umpteen network calls and file I/O there is very little for which the act of measuring will in any way significantly degrade the performance of the operation being measured.
Of course if you write an instrumentation message to a log file for every disk I/O request you’ll probably spend more time measuring than doing useful work and so the level of measurement needs to be granular enough to capture something useful without being too intrusive or verbose. This may itself depend on how and where you are capturing your data as you probably don’t want to be contending with the resource that you’re trying to measure.
As a starting point I would expect to measure every call out to a remote service, whether that is a web service, database query, etc. Every remote call has so many other moving parts in play that it’s almost certainly going to act up at some point. Reading and writing files to disk, especially when that disk is based remotely on a NAS/SAN/DFS is also a must, although the granularity would be to read/write the entire file. On top of those any intensive memory or CPU hogs should also get instrumented as a matter of course.
There is one scenario when I would consider discarding the measurements for a task and that’s when an exception has been thrown. If an error has occurred you cannot tell how much of the operation was performed, which for a lost connection could be none, and so having lots of dubious tiny points in your data will only succeed in distorting it.
Types of instrument
You only need a few basic instruments to help you capture most of the key metrics. Some of those instruments are ‘real’, whereas others may need to be synthesized by using API’s from the OS or language runtime.
Wall clock
By far the simplest instrument is the good old fashioned clock. The term ‘wall clock’ is often used to distinguish it from other internal clocks as it measures elapsed time as perceived by the user (or operator). These are probably the easiest to implement as virtually every runtime has the ability to report the current date and time, and so therefore you can calculate the difference.
The main thing you need to watch out for is the difference between the precision and the accuracy of the clock. Although you may be able to format the date and time down to 1 nanosecond that doesn’t mean it’s captured at that resolution. For example the .Net
System.DateTime type has a precision measured in nanoseconds but is only accurate to around 16 ms [Lippert]. As things stand today I have found this good enough to capture the vast majority of ‘big picture’ measurements.
There are often other platform specific alternatives to the classic wall clock, some are old fashioned like the original
GetTickCount() API in Windows which also only has a resolution of 10–16 ms on the NT lineage (it was a whopping 50 ms under 3.x/95). The Windows
QueryPerformanceCounter() API in contrast has a much, much higher resolution, due to the hardware used to derive it, and it’s from this that the .Net
StopWatch type gains its sub-microsecond precision and accuracy [Lippert]. However, be warned that this impressive feat is heavily dependent on the hardware and is therefore more usable server-side than client-side.
Specialised clocks
Internally operating systems often track metrics around process and thread scheduling. One such bit of data Windows provides is the amount of time spent in the OS kernel as opposed to user space. Coupled with the elapsed time you can discover some useful facts about your process.
If it’s a high-performance computing (HPC) style engine that is expecting to be CPU bound, you might expect all its time to be spent in user space, but if it’s in neither you’re probably doing more I/O than you expect. I have observed a high amount of time in the kernel when there are a large number of exceptions bouncing around (and being ignored) or when the process is doing an excessive amount of logging. The converse can also be true about an I/O bound task that seems to be consuming too much CPU, such as during marshalling.
Network latency
Prior to Windows 2000 the clocks on Windows servers were notoriously bad, but with the addition of support for Kerberos the situation got much better (it had to) as servers now synchronised their clocks with a domain controller. With servers having a much better idea of what the current time is, you can begin to capture the network latency incurred by a request – this is the time it takes for the request to be sent and for the response to the received.
You can measure the latency by timing the operation from the client’s perspective (the remote end, server-wise) and also within the service handling the request (the local end, server-wise) and then calculating the difference. For example, if the client sees the request taking 10 seconds and the service knows it only spent 8 seconds processing it, the other 2 seconds must be spent in and around the infrastructure used to transport the request and response. Although you might not be able to correctly apportion the time to the request or response sides (as that relies on the absolute times) you can at least calculate the total latency which only relies on the relative differences.
This kind of metric is particularly useful anywhere you have a queue (or non-trivial buffering) in place as it will tell you how long a message has been stuck in the pending state. For example in a grid computing scenario there is usually a lot of infrastructure between your client code and your service code (which is executed by a remote engine process) that will queue your work and marshal any data to and from the client. However, if the queue has the ability to retry operations you will have to allow for the fact that the latency might also include numerous attempts. This is one reason why handling failures and retries client-side can be beneficial.
Memory footprint
Aside from the use or idleness of your CPUs, you’ll also want to keep track of the memory consumed by your processes. Sadly this is probably not tracked at a thread level like CPU times because a global heap is in use. However, single-threaded processes are often used for simplicity and resilience and so you might be in a position to track this accurately in some parts of the system.
On Windows you generally have to worry about two factors – virtual address consumption and committed pages. Due to the architecture of Windows’ heaps you can end up in the mysterious position of appearing to have oodles of free memory and yet still receive an out-of-memory condition. See my previous Overload article on breaking the 4 GB limit with 32-bit Windows processes for more details [Memory]. Suffice to say that process recycling is a common technique used with both managed and native code to work around a variety of heap problems.
For services with some form of caching built-in memory footprint metrics will help you discover if your cache eviction policy is working efficiently or not.
Custom measurements
One of the problems with any measurement taken in isolation is that you often need a little more knowledge about what the ‘size’ of the task actually was. If your workload varies greatly then you’ll not know if 1 second is fast, acceptable or excessive unless you also know that it had thousands of items to process. Then when you’re analysing your data you can scale it appropriately to account for the workload.
To help provide some context to the instrument measurements it’s worth accumulating some custom measurements about the task. These might include the number of items in the collection you’re reading/writing or the number of bytes you’re sending/receiving.
Causality
Every similar set of measurements you capture need to be related in some way to an operation and so you also need to capture something about what that is and also, if possible, some additional context about the instance of that operation. Once you start slicing and dicing your data you’ll want to group together the measurements for the same operation so that you can see how it’s behaving over time.
Additionally when you start to detect anomalies in your systems’ behaviour you’ll likely want to investigate them further to see what led to the scenario. Whilst the date and time the data was captured is a good start, it helps if have some kind of ‘transaction identifier’ that you can use to correlate events across the whole system. I described one such approach myself in a previous edition of Overload [Causality].
Output sinks
Once you’ve started accumulating all this data you’ll want to persist it so that you can chew over it later. There are various options, none of which are mutually exclusive, and so it might make sense to publish to multiple targets for different reasons – usually based around active monitoring or passive offline investigation.
One thing you’ll definitely want to bear in mind is how you handle errors when writing to an output sink fails. You almost certainly don’t want the failure to cause the underlying operation to also fail. If the sink is remote you could try buffering the data for a short while, but then you run the risk of destabilising the process, and eventually the machine [Memory], if you buffer too much for too long.
Log file
The humble (text based) log file is still with us, and for good reason. Many of us have honed our skills at manipulating these kinds of files and as long as you keep the format simple, they can provide a very effective medium.
The single biggest benefit I see in storing your instrumentation data in your log files is the locality of reference – the data about the operation’s performance is right there alongside the diagnostic data about what it was doing. When it comes to support queries you’ve already got lots of useful information in one place.
One method to help with singling out the data is to use a custom severity level, such as ‘PRF’ (for performance), as this will use one of the key fields to reduce the need for any complex pattern matching in the initial filtering. The operation itself will need a unique name, which could be derived from the calling method’s name, so that you can aggregate values. And then you’ll need to the data encoded in a simple way, such as key value pairs. Here is an example log message using that format:
2013-06-26 17:54:05.345 PRF [Fetch Customers] Customer:Count=1000;WallClock:DurationInMs=123;
Here we have the operation name –
Fetch Customers – and two measurements – the number of items returned followed by the elapsed time in milliseconds. Although the simpler key names
Count and
Duration may appear to suffice at first, it’s wise to qualify them with a namespace as you may have more than one value captured with the same name. Including the units also helps if you capture the same value in two different formats (e.g. milliseconds and
TimeSpan). This might seem redundant, but when reading a log for support reasons it helps if you don’t have to keep doing arithmetic to work out whether the values are significant or not.
Classic database
Where better to store data than a database? Your system may already capture some significant timings in your main database about the work it’s doing, such as when a job was submitted, processed and completed. And so you might also want your instrumentation data captured alongside it to be in one place. Some people are more comfortable analysing data using SQL or by pulling portions of data into a spreadsheet as not everyone is blessed with elite command line skills and intricate knowledge of the Gnu toolset.
Of course when I say ‘classic’ database I don’t just mean your classic Enterprise-y RDBMS – the vast array of NOSQL options available today are equally valid and quite possibly a better fit.
Round Robin Database
For analysis of long term trends the previous two options probably provide the best storage, but when it comes to the system’s dashboard you’ll only be interested in a much shorter time window, the last 10 minutes for example. Here you’ll want something akin to a fixed-size circular buffer where older data, which is of less interest, is either overwritten or pruned to make room for newer values.
There are a few variations on this idea; one in particular is the Round Robin Database (or RRDtool [RRDtool]). This prunes data by using a consolidation function to produce aggregate values that lower the resolution of older data whilst still maintaining some useful data points.
Performance counters
Another variation of the circular buffer idea is to expose counters from the process itself. Here the process is responsible for capturing and aggregating data on a variety of topics and remote processes collect them directly instead. This has the advantage of not requiring storage when you don’t need it as you then log only what you need via a remote collector. However, back at the beginning I suggested you probably want to capture everything you can afford to, so this benefit seems marginal.
One word of caution about the Windows performance counter API – it’s not for the faint of heart. It is possible that it’s got better with recent editions (or via a 3rd party framework) but there are some interesting ‘issues’ that lie in wait for the unwary. That said the concept still remains valid, even if it’s implemented another way, such as using a message bus to broadcast a status packet.
A simple instrumentation framework
Anyway, enough about the theory what does this look like in code. You shouldn’t be surprised to learn it’s not rocket science. Basically you just want to start one or more instruments before a piece of functionality is invoked, and stop them afterwards. Then write the results to one or more sinks.
If you’re working with C++ then RAII will no doubt have already featured in your thoughts, and for C# developers the Dispose pattern can give a similar low noise solution. The effect we want to achieve looks something like Listing 1 (I’m going to use C# for my examples).
This code, which is largely boilerplate, adds a lot of noise to the proceedings that obscures the business logic and so we should be able to introduce some sort of façade to hide it. What I’d be looking for is something a little more like Listing 2.
Hopefully this looks far less invasive. Through the use of reflection in C# (or pre-processor macros in C++) we should be able to determine a suitable name for the operation using the method name.
By switching to flags for the instrument types, we’ve factored out some noise but made customising the instruments harder. Luckily customisation is pretty rare and so it’s unlikely to be a problem in practice. The façade could also provide some common helper methods to simplify things further by wrapping up the common use cases, for example:
using (MeasureScope.WithWallClock())
or
using (Measure.CpuBoundTask())
The one other minor detail that I have so far glossed over is how the output sink is determined. For cases where it needs to be customised it can be passed in from above, but I’ve yet to ever need to do that in practice. Consequently I tend to configure it in
main() and store it globally, then reach out to get it from behind the façade.
The façade
The façade used by the client code can be implemented like Listing 3.
Retrieving the calling method name in C# (prior to C# 4.0) relies on walking the stack back one frame. Of course if you write any wrapper methods you’ll need to put the stack walking in there as well or you’ll report the wrapper method name instead of the business logic method name (see Listing 4).
The same rule applies if you’ve used closures to pass your invocation along to some internal mechanism that hides grungy infrastructure code, such as your Big Outer Try Block [Errors]. See Listing 5.
Bitwise flags seem to be a bit ‘old school’ these days, but for a short simple set like this they seem about right, plus you have the ability to create aliases for common permutations (see Listing 6).
The instrument controller
The behind-the-scenes controller class that starts/stops the instruments and invokes the outputting is only a tiny bit more complicated than the façade (Listing 7).
The sink
The measurements that are returned from the instruments are just string key/value pairs. Given their different nature – datetime, timespan, scalar, etc. – it seems the most flexible storage format. I’ve implemented the sink in Listing 8 as a simple wrapper that writes to a log file through some unspecified logging framework.
Here I’ve assumed that the Log façade will not allow an exception to propagate outside itself. That just leaves an out-of-memory condition as the only other sort of failure that might be generated and I’d let that propagate normally as it means the process will likely be unstable.
The instruments are equally simple – a pair of methods to control it and a property to retrieve the resulting metric(s). Like all ‘good’ authors I’ve chosen to elide some error handling in the
Start() and
Stop() methods to keep the code simple (Listing 9), but suffice to say you probably want to be considering what happens should they be called out of sequence.
Earlier I mentioned that I would avoid writing measurements taken whilst an exception was being handled; that can be achieved in C# with the following check [Exception]. See Listing 10.
Summary
This article has shown that it’s relatively simple to add some instrumentation calls within your own code to allow you to track how it’s performing on a continual basis. It explained what sort of measurements you might like to take to gain some useful insights into how your service calls and computations are behaving. It then followed up by showing how a simple framework can be written in C# that requires only out-of-the box techniques and yet still remains fairly unobtrusive.
Once in place (I recommend doing this as early as possible, as in right from the very start) you’ll be much better prepared for when the inevitable happens and the dashboard lights up like a Christmas tree and the phones start buzzing.
Acknowledgements
Thanks to Tim Barrass for introducing me to the Round Robin Database as a technique for persisting a sliding window of measurements and generally helping me refine this idea. Additional thanks as always to the Overload advisors for sparing my blushes by casting their critical eye and spotting my mistakes.
References
[Causality] Chris Oldwood, ‘Causality – relating distributed diagnostic contexts’, Overload 114
[Errors] Andy Longshaw and Eoin Woods, ‘The Generation, Management and Handling of Errors (Part 2)’, Overload 93
[Exception] Ricci Gian Maria, ‘detecting if finally block is executing for an manhandled exception’,
[Hoare]
[Lippert] Eric Lippert, ‘Precision and accuracy of DateTime’,
[Memory] Chris Oldwood, ‘Utilising more than 4GB of memory in 32-bit Windows process’, Overload 113
[RRDtool] | https://accu.org/index.php/journals/1843 | CC-MAIN-2018-47 | refinedweb | 3,750 | 54.86 |
.
@luz I don't understand the problem. GPIO18 and GPIO19 work fine with my latest change, I can see the pin-state change when I connect the pin to HIGH or LOW.
@Lazar-Demin said in How to switch AGPIO (GPIO 18/19) - Omega2 FW vs LEDE build:
Also, we've open-sourced
omega2-ctrl, check it out here:
Great! Thanks a lot!
This allowed me to figure out what the "ephy" setting that confused me actually is.
It controls EPHY_LED0_N_JTDO pin to be either GPIO43 or the Ethernet LED for Port 0 (P0_LED_AN_MODE, Bits 3..2 in GPIO2_MODE). Thus, it should probably be called "eled". On the other hand, as that pin seems not connected in the Omega2, maybe it should be removed from omega2-ctrl, together with "wled" which seems not connected, too.
@WereCatf said in How to switch AGPIO (GPIO 18/19) - Omega2 FW vs LEDE build:
@luz I don't understand the problem. GPIO18 and GPIO19 work fine with my latest change, I can see the pin-state change when I connect the pin to HIGH or LOW.
It's definitely not a .dts thing. If I manually disable the AGPIO (see devmem command in the original post), GPIO18/19 work fine.
But some software piece in my LEDE build does switch this entire pin group (all 16 bins related to ethernet switch port 1..4) into AGPIO/ethernet mode.
In the meantime, I have a suspect: it could be the mere presence of the "swconfig" package. This is a utility to control built-in switch hardware, but it also pulls in a kernel module and maybe this module enables the switch hardware. I'm now building a new image without swconfig, let's see if that helps...
Could you please check in your .config, do you have "swconfig" package selected?
@luz I have
CONFIG_PACKAGE_swconfig=ybut
CONFIG_PACKAGE_kmod-swconfig=m-- perhaps that's the difference? You could also try if just disabling the switch in the DTS-file works, e.g. add the following to the DTS:
&esw { status = "disabled"; };
@WereCatf disabling the switch in the device tree kills ethernet functionality entirely :-(
So this is no option for me - I need the ethernet port.
Maybe its using the ethernet port that causes AGPIO to get enabled (and GPIO18/19 disabled)?
Do you have ethernet in use on your omega2?
BTW: CONFIG_PACKAGE_swconfig=y and CONFIG_PACKAGE_kmod-swconfig=m didn't help either.
@WereCatf I mean, is the eth0 interface enabled (in software)? It does not matter whether you have a ethernet dock or not.
But if it isn't enabled in your configuration, then that could be the reason why GPIO18/19 work for you, and don't work in my case.
Hi,
Yeah, probably I'm a thread necromancer... But I think I found why (when, how) the official Omega2 image sets AGPIO to digital IO, and which is not done by the original LEDE project.
I did the following trick, and it seems it succeeded: I connect one channel of my scope to one of the GPIOs (
GPIO17) which was ~1.65V by default during boot and changed to 0V at a point in time and I attached the other probe to the TX line.
As it clearly show the next screenshot, the line switched to digital at SD card initialization:
It changes right after the kernel prints
MSDC device init.. So, I searched for it in the kernel source and I found.
Here is the corresponding snippet (it is located in
linux-4.9.37/drivers/mmc/host/mtk-mmc/sd.c):
static int __init mt_msdc_init(void) { int ret; /* +++ by chhung */ u32 reg; #if defined (CONFIG_MTD_ANY_RALINK) extern int ra_check_flash_type(void); if(ra_check_flash_type() == 2) { /* NAND */ printk("%s: !!!!! SDXC Module Initialize Fail !!!!!", __func__); return 0; } #endif printk("MTK MSDC device init.\n"); mtk_sd_device.dev.platform_data = &msdclinux-4.9.37/drivers/mmc/host/mtk-mmc/sd.c0_hw; if (ralink_soc == MT762X_SOC_MT7620A || ralink_soc == MT762X_SOC_MT7621AT) { //#if defined (CONFIG_RALINK_MT7620) || defined (CONFIG_RALINK_MT7621) reg = sdr_read32((volatile u32*)(RALINK_SYSCTL_BASE + 0x60)) & ~(0x3<<18); //#if defined (CONFIG_RALINK_MT7620) if (ralink_soc == MT762X_SOC_MT7620A) reg |= 0x1<<18; //#endif } else { //#elif defined (CONFIG_RALINK_MT7628) /* TODO: maybe omitted when RAether already toggle AGPIO_CFG */ reg = sdr_read32((volatile u32*)(RALINK_SYSCTL_BASE + 0x3c)); reg |= 0x1e << 16; sdr_write32((volatile u32*)(RALINK_SYSCTL_BASE + 0x3c), reg); reg = sdr_read32((volatile u32*)(RALINK_SYSCTL_BASE + 0x60)) & ~(0x3<<10); #if defined (CONFIG_MTK_MMC_EMMC_8BIT) reg |= 0x3<<26 | 0x3<<28 | 0x3<<30; msdc0_hw.data_pins = 8, #endif //#endif } sdr_write32((volatile u32*)(RALINK_SYSCTL_BASE + 0x60), reg); //platform_device_register(&mtk_sd_device); /* end of +++ */ ret = platform_driver_register(&mt_msdc_driver); if (ret) { printk(KERN_ERR DRV_NAME ": Can't register driver"); return ret; } printk(KERN_INFO DRV_NAME ": MediaTek MT6575 MSDC Driver\n"); #if defined (MT6575_SD_DEBUG) msdc_debug_proc_init(); #endif return 0; }
It contains the part which actually do the job:
reg = sdr_read32((volatile u32*)(RALINK_SYSCTL_BASE + 0x3c)); reg |= 0x1e << 16; sdr_write32((volatile u32*)(RALINK_SYSCTL_BASE + 0x3c), reg);
It sets all the four bits of
AGPIO_CFG(
EPHY_GPIO_AIO_EN, bit 20..17 (
GPIO14,
GPIO15,
GPIO16and
GPIO17)) to
1- thus enabling them as digital GPIO.
So, it does not happen in any user space utility (like
swconfig), but it happens in the kernel by enabling SD card.
Hope this is still useful for some of you ;-)
/sza2
@sza2-sza2 wow! Impressive analysis and I admire your patience to figure that out!
Indeed, that is useful to know. Technically I had „solved“ the problem with directly writing to the AGPIO_CFG register, but I was wary of that hack without knowing why this was needed in the first place. Now I feel better ;-)
LEDE is a router box OS, so it makes sense it makes the choice for Ethernet (AGPIO) by default. But still, it has to make that choice at some point, because after reset, the MT7688 is not in AGPIO mode. So there must be code somewhere else which enabkes AGPIO mode on in the first place. Probably low level ethernet init code...
IMHO it would make sense for a non-router device like the Omega2 to never ever switch these pins to AGPIO, because otherwise there's a short period of unpredictable signal level left between reset and the time whatever code has a chance to switch back to GPIO mode, making them unusable for some purposes.
Your work motivates me to dig in a bit and find (and eventually patch) that initialisation code :-)
I cannot recall perfectly, but as far as I remember, those pins became ~1.65V during bootloader time. Actually, I saw a glitch on the line somewhere before it permanently set to 0V, and it was already during kernel boot (maybe around MTD initialization) but I did not dig deeper. Probably, later on I'll check it - although my online radio receiver already works, so there is no compelling reason :-)
/sza2 | https://community.onion.io/topic/1266/how-to-switch-agpio-gpio-18-19-omega2-fw-vs-lede-build | CC-MAIN-2019-26 | refinedweb | 1,104 | 61.67 |
Re: WaitForSingleObject() will not deadlock
- From: Joseph M. Newcomer <newcomer@xxxxxxxxxxxx>
- Date: Mon, 02 Jul 2007 03:01:35 -0400
And in what way am I wrong about Windows mutexes?
I have been doing concurrent programming professionally for 32 years in a variety of
operating systems and environments, and I have considerable expertise in it. I suppose
you're next going to say years of experience don't qualify me either. On the other hand,
I was absolutely right about the semantics of mutexes (since I can give you quotes from
the documentation which are consistent with my statements of how mutexes work) and you
were wrong (because you said I didn't understand mutexes, even though I told you your
expectations would violate the *documented* mutex behavior). I started working with
concurrency shortly after Dijkstra's 1968 paper, but only worried about it between about
1968 and 1972, building only experimental systems (my first synchronization system was
written in 1968 as an interterminal chat system running on TSS/360, so as someone who has
actually written synchronization code, I've now been doing it for 39 years, with
considerable success, I might add). I wrote the P and V operations for our student
project on the IBM/360 computer. Between 1975 and 1977 I did a substantial amount of work
on a 16-processor multiprocessor system, and our problems included concurrent update of
objects, deadlock avoidance, and lock contention avoidance. I spent a fair amount of time
developing algorithms for analyzing deadlock in source code by doing static source
analysis and using Petri net simulations. The C.mmp/Hydra system (C.mmp was the
multiprocessor system, built from 16 PDP-11 computers, and Hydra was the capability-based
object-oriented operating system that ran on it) was one of the earliest mid-scale
multiprocessor systems. I was part of the Ada evaluation team and studied the Ada
rendezvous mechanism in some detail including the issues of implementing it, although we
did not publish any papers on this aspect.
Whatever you care to believe, I was right about mutexes. You were not.
More below...
*****
On Sun, 01 Jul 2007 18:45:05 -0700, Frank Cusack <fcusack@xxxxxxxxxxx> wrote:
On Sun, 01 Jul 2007 16:08:59 -0400 Joseph M. Newcomer <newcomer@xxxxxxxxxxxx> wrote:*****
Two weeks ago, I taught my Systems Programming Course, in which we
use mutexes in the lab on Thursday. I also teach about critical
sections. Last week I taught our Device Driver course, in which I
taught what a mutex is. As well as spin locks, fast mutexes, and
executive resources. I suspect I actually understand what a mutex
is.
Teaching a subject does not qualify you as an expert in it. I have
had some very qualified and very good teachers, but like all of us I
have also had some poor ones. I'm not outright saying you are
unqualified, and I am not implying anything about your teaching skills
(after all, I have no idea), I'm simply saying that your assertion of
qualification or expert knowledge by teaching about it is meaningless.
Furthermore, unless you are talking about a 40 hour course, covering
the topic of synchronization primitives or device drivers in "a week"
is just introductory and again does not demonstrate expert knowledge.
For example, you asked a question which clearly shows you do not
undertand recursive acquisition semantics, one of the fundamental
properties of a mutex. I quote from your post:
if (WaitForSingleObject(mutex, INFINITE) == WAIT_OBJECT_0)
AfxMessageBox("acquired", ID_OK);
if (WaitForSingleObject(mutex, INFINITE) == WAIT_OBJECT_0)
AfxMessageBox("acquired", ID_OK);
You got two message boxes, which is EXACTLY what you should see happen!
Now, if I say :"this would violate the fundamental design of a mutex
to have this deadlock", and it is obvious that it does not deadlock,
it is almost certainly because I understand how mutexes work. You
are insisting that it *should* deadlock, and this shows you do *not*
understand how a mutex works.
Get it straight. I did not insist that it should deadlock, I asked
why it didn't. I didn't find documentation that Windows mutexes were
recursive, so I had to ask. And it's not obvious why it didn't
deadlock; there are 2 possible reasons.
1. The mutex is recursive.
2. WaitForSingleObject() knows that the object has already been
"signalled", and therefore returns success, WITHOUT again
locking the mutex.
How is (2) different from (1)? Note that once a mutex is non-signaled, it is owned, and
would not be "locked" again. Instead, the reference count is is incremented, as is done
with all implementations of recursive mutexes I know of. Furthermore, the documentation
actually states that mutexes can be acquired recursively because it states that a WaitFor
will not block, and as many ReleaseMutex calls as successful WaitFor calls must be
performed.
*****
**********
As a newcomer to Windows, the WaitForSingleObject() call is confusing
and has unclear semantics to me. In fact, the documentation for
CreateMutex() talks about the mutex being "signaled" and "unsignaled"
as opposed to acquired and released. That was confusing for me.
Hell, I had a hard time figuring out that the '0' at the end of
WAIT_OBJECT_0 was a '0' and not an 'O'.
Objects are referred to as being signaled or non-signaled. Mutexes are a subset of the
general waitable object model, and therefore are either signaled or non-signaled just like
any other waitable object. Acquiring a mutex is a mutex-specific operation that converts
the mutex object from the signaled state to the non-signaled state, so it would not be
correct to talk about acquisition and release since that would limit the discussion to a
single type of object, and would not be applicable to file handles, waitable timers,
events, change notifications, job objects, processes, threads, or semaphores, none of
which have a concept of "acquisition". The only unifying concept is "signaled" and
"nonsignaled", and mutex acquisition is an operation that changes the state of a
particular kind of the waitable objects, a mutex, from signaled to non-signaled.
******
********
What you are asking violates the fundamental design of a mutex. I
Fundamental? ha. You mean fundamental to the design of a Windows
mutex.
This *is* a Windows newsgroup, and the behavior I described *is* fundamental to the design
of mutexes in Windows, so what is wrong with what I said? Your desires violate the
fundamental design of a mutex. In the context of a Windows newsgroup, that is the only
mutex that would be discussed.
****
**********
would suggest that you are the one who doesn't understand them. Of
I agree, I do not understand Windows mutexes. Well, now I understand
them a bit more.
So why did you presume I didn't understand them? I've been using synchronization
primitives for nearly 40 years, starting with the test-and-set instruction of the IBM 360
back in 1968.
*****
**********
course, had you actually read the documentation on a mutex, you
would have seen the paragraph that says (and I have pasted this from
the MSDN documetation on CreateMutex):
================================
."
=================================
Thanks, I see that now. Wish I had seen it earlier, would have
answered my question right away.
So you are asking how to create a situation that violates the
fundamental *documented* design of a mutex, and I told you that your
suggestion violates the fundamental design of a mutex; your response
is that I have a limited understanding of what a mutex is, so please
explain exactly what part of a mutex I don't understand?
I will agree that you understand what a Windows mutex is (expert
knowledge even), but in my mind the jury is still out on whether or
not you understand more fundamentally about mutexes in general. Not
saying you don't have a solid grasp, just that your assumptions are
Windows-oriented.
Gee, after 39 years of writing synchronization code, I probably have discovered SOMETHING
about synchronization, the basic principles,the high-level models, the formal models, and
a lot of other details. But when a question appears in a Windows group, showing an
example using the Windows API, I tend to assume that we are talking about Windows mutexes,
and not some general model of mutexes through history. If you ask a question about the
presidential candidate in a Republican blog, it is unlikely you are talking about the
office as it was in 1780, or about the presidential candidate of some South Pacific island
or of some company or social organization. So there is no reason to assume that we are
talking about POSIX semantics, or Dijkstra's P/V primitives, or Hoare monitors, or
discussing the Multics wait/wake problem, or Brinch-Hansens signaling mechanism.
With respect to my background, Per Brinch-Hansen's office was across the hall from mine;
one of Dijkstra's best students, Nico Habermann, was a professor at CMU and eventually
department head, and he taught the operating systems course there, which I sat in on. My
Software Systems PhD qualifier exam, which had several questions on synchronization, was
complimented by Nico as being the best answer he'd ever seen on the subject. I learned
initially about these issues from Dave Parnas, one of the creators of modular programming,
if not THE creator of modular programming. Have you read Jerry Salzer's work on
synchronization issues in Multics? Our first student exercise in operating systems was to
create P and V using ordinary instructions with the possiblity of context swap caused by
thread preemption between any two instructions, and the synchronization had to still work.
Let's see, have I amply demonstrated that I actually DO understand synchronization
primitives in general? But if you ask a question in a Windows newsgroup, expect to get a
Windows-oriented answer. If you need a general answer, go find some general newsgroup,
comp.something-dealing-with-synchronization. If you ask a Windows question in a Windows
newsgroup, expect a Windows-oriented answer.
But the other basic issue was that you were trying to use a synchronization primitive to
solve a non-synchronization problem, and the solution would have failed no matter what the
semantics were. With recursive acquisition, there is no synchronization, and without it,
the one-and-only GUI thread deadlocks itself, so both possible implementations of
synchronization primitives, the results are incorrect. You didn't have a synchronization
problem in the first place, just a problem of doing the same thing twice, in the same
thread.
******
**********
I will submit that it was a mistake on my part to consider any mention
of a mutex in this newsgroup as meaning anything other than a Windows
mutex. Double dumbass on me.
Can you explain and/or point me to documentation on the memory
visiblity semantics of a Windows mutex? I am not trying to draw you
out; this is an honest question.
What's "memory visibility semantics"? I've not heard that term applied to any
synchronization object, but if you explain what you mean, I can probably give you an
answer.
I tried a google search on "memory visibility semantics mutex" but didn't find anything
that seemed relevant.
Mutexes are not memory objects; they are represented by handles. So there is no memory to
be visible. But the handle can be visible across processes, and there are some
interesting protocols, such as using GUIDs for names, to prevent collision of these names
in the global namespace.
You cannot share a handle by placing the handle in shared memory, because the handle would
not be in the process handle table of the other processes. The usual technique of making
a mutex, semaphore, or event visible to another process is to have each process do
HANDLE mutex = CreateWhatever(NULL, otherparametershereasappropriate, MY_OBJECT_NAME);
where you do something like
#define MY_OBJECT_NAME _T("MyQueueMutex-{8D9BC3AE-A745-4513-B4D1-4AC600E62888}")
where you use the program GUIDGEN.EXE to create that GUID. This guarantees that no one,
anywhere, ever, will accidentally create a mutex of the same name you have. Actually, the
GUID is sufficient, but you use the readable name to communicate information to the person
using it (there are tools that can inspect the object space, for example).
The effect is to allow each process to create a new handle to a mutex. If the mutex is
named, and already exists, CreateMutex == OpenMutex. This means that it doesn't matter
what order the processes start up in; one gets to create the mutex and the rest get to use
it.
To use a handle, it must be in the handle table of the process. There are several ways
get a handle there, such as doing CreateMutex or the appropriate action for the type of
object whose handle you want, creating a duplicate handle using DuplicateHandle and
targeting the receiving process, or creating an inheritable handle which is inherited by a
child process. However, far and away the simplest method for synchronization objects is
to simply do the appropriate Create in the processes that wish to share.The other
techniques have problems in communicating the numerical value of the handle to the target,
and offer no serious advantages over the Create method.
joe
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Opened 10 years ago
Closed 10 years ago
#7424 closed (wontfix)
Near mentions of TEMPLATE_DIRS absolute path, include tip dynamic absolute path
Description
Most documentation of settings.TEMPLATE_DIRS instructs the user to use absolute paths, in the form of a hardcoded string literal absolute path.
A new user would immediately wonder if they needed to plan their deployment absolute path right then and there.
To avoid this surprise, it would be helpful to note that the absolute path can be created dynamically, such as with:
TEMPLATE_DIRS = (os.path.join(os.path.abspath(os.path.dirname(__file__)), 'templates'),)
Change History (2)
comment:1 Changed 10 years ago by
comment:2 Changed 10 years ago by
Listing everything you might do in this setting (which is the road this is heading down) isn't really efficient or useful. Also, re-using a settings file without changes in multiple environments isn't necessarily a design goal -- settings will always, at least to some extent, be specific to the environment in which Django is being deployed.
+1 from me. The absolute path is useful in other places as well (DATABASE_NAME for sqlite and MEDIA_ROOT), so I personally always put
{{
import os
PROJDIR = os.path.dirname(os.path.abspath(file))
}}
into
settings.pyand use PROJDIR where absolute path is required.
Marking as accepted as projects can not be really used in different environments without this. | https://code.djangoproject.com/ticket/7424 | CC-MAIN-2018-39 | refinedweb | 230 | 52.29 |
One of the truly nice things about blogging every day is that often I find my next topic suggested and even outlined by folks that are crazy enough to be reading along. One such individual, Robert, suggested in the comments of last night's post that I try a different approach with Dart
pub deploycommand.
As and aside, comments and suggestions are especially welcome and appreciated when I have spent too much of the day enjoying a conference and do not have the proper amount of time to plan (or think) about a post. And yes, contrary to public perception, I do put some actual thought into these things beforehand. I wonder why I am rambling...
Anyhow, the goal remains the same: to see if I can get the
pub deploycommand to deploy my code from this state:
➜ ice-beta git:(gh-pages) ✗ tree . -P 'index*|*.dart' -l -L 2 . ├── index.html ├── main.dart └── packages ├── browser -> /home/chris/.pub-cache/hosted/pub.dartlang.org/browser-0.6.5/lib ├── crypto -> /home/chris/.pub-cache/hosted/pub.dartlang.org/crypto-0.6.5/lib ├── ice_code_editor -> /home/chris/.pub-cache/hosted/pub.dartlang.org/ice_code_editor-0.0.9/lib ├── js -> /home/chris/.pub-cache/hosted/pub.dartlang.org/js-0.0.24/lib ├── meta -> /home/chris/.pub-cache/hosted/pub.dartlang.org/meta-0.6.5/lib └── unittest -> /home/chris/.pub-cache/hosted/pub.dartlang.org/unittest-0.6.5/lib 7 directories, 2 filesInto something that will work on GitHub pages. The primary obstacle to this is the symbolic links that Dart Pub normally uses. The other obstacle being that the
pub deploycommand seems to expect code to be in a web sub-directory.
Robert's suggestion was to place a main entry point in the web sub-directory that imports the real (i.e. the current) main.dart entry point. The hope being that this would let the code remain as-is, but still allow
pub deployto to the necessary compilation to JavaScript and removal of the packages symbolic links.
So I create
web/main.dartas:
import '../main.dart' as Main; main()=> Main.main();Then try a
pub deploy:
➜ ice-beta git:(gh-pages) ✗ pub deploy Finding entrypoints... Copying web/ → deploy/ Compiling web/main.dart → deploy/main.dart.js web/main.dart:1:8: Error: no library name found in import '../main.dart' as Main; ^^^^^^^^^^^^^^ Failed to compile "/home/chris/repos/gamingjs/ice-beta/web/main.dart".So I add a library statement to the “real” main script:
library main; import 'package:ice_code_editor/ice.dart' as ICE; main()=> new ICE.Full();With that
pub deployworks:
➜ ice-beta git:(gh-pages) ✗ pub deploy Finding entrypoints... Copying web/ → deploy/ Compiling web/main.dart → deploy/main.dart.jsBut the layout still does not seem as though it is going to work. The
index.htmlin my sub-directory points to the main.dart file in my sub-directory. For Dart-enabled browsers that will work just fine, except that the symbolic links will still not resolve once published to GitHub pages:
➜ ice-beta git:(gh-pages) ✗ tree . -P 'main.dart.js|index*|*.dart' -l -L 2 . ├── deploy │ ├── main.dart │ └── main.dart.js ├── index.html ├── main.dart ├── packages │ ├── browser -> /home/chris/.pub-cache/hosted/pub.dartlang.org/browser-0.6.5/lib │ ├── crypto -> /home/chris/.pub-cache/hosted/pub.dartlang.org/crypto-0.6.5/lib │ ├── ice_code_editor -> /home/chris/.pub-cache/hosted/pub.dartlang.org/ice_code_editor-0.0.9/lib │ ├── js -> /home/chris/.pub-cache/hosted/pub.dartlang.org/js-0.0.24/lib │ ├── meta -> /home/chris/.pub-cache/hosted/pub.dartlang.org/meta-0.6.5/lib │ └── unittest -> /home/chris/.pub-cache/hosted/pub.dartlang.org/unittest-0.6.5/lib └── web ├── main.dart └── packages -> ../packages [recursive, not followed] 10 directories, 5 filesFurther, the index.html and main.dart in my web application directory will not work with non-Dart browsers since there is no
main.dart.jsfile in there for legacy browser to fallback to.
I might be able to post-process this with another script—moving the deploy/main.dart.js into the application directory. But I already have a decache.sh script that solves my current use-case. So unless I am missing something (which is certainly possible in my current state—any ideas Robert?) it's back to OSCON partying. Woo hoo!
UPDATE: Per a suggestion from Paul Evans, I tried setting the
HOMEshell variable to the current working directory:
➜ ice-beta git:(gh-pages) ✗ HOME=`pwd` pub install Resolving dependencies..................... Downloading js 0.0.24 from hosted... Downloading ice_code_editor 0.0.9 from hosted... Downloading browser 0.6.5 from hosted... Downloading meta 0.6.5 from hosted... Downloading crypto 0.6.5 from hosted... Downloading unittest 0.6.5 from hosted... Dependencies installed!(using
HOME=.creates packages symlinks that do not resolve)
That looks promising in that all of my dependencies, which would otherwise be re-used from the system cache, are downloaded again. However, I am still left with a symbolic links to a package cache—only the package cache is now in the current directory:
➜ ice-beta git:(gh-pages) ✗ ls -l packages total 4 lrwxrwxrwx 1 chris chris 54 Jul 26 11:16 browser -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/browser-0.6.5/lib lrwxrwxrwx 1 chris chris 53 Jul 26 11:17 crypto -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/crypto-0.6.5/lib lrwxrwxrwx 1 chris chris 62 Jul 26 11:16 ice_code_editor -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/ice_code_editor-0.0.9/lib lrwxrwxrwx 1 chris chris 50 Jul 26 11:17 js -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/js-0.0.24/lib lrwxrwxrwx 1 chris chris 51 Jul 26 11:16 meta -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/meta-0.6.5/lib lrwxrwxrwx 1 chris chris 55 Jul 26 11:16 unittest -> /home/chris/repos/gamingjs/ice-beta/.pub-cache/hosted/pub.dartlang.org/unittest-0.6.5/libThat is not really an improvement. The
/home/chrisdirectory does not exist on GitHub. Even if it did, the symbolic links will not work in GitHub pages. Even if they did, “dot” files and directories like
.pub-cachewill not work with GitHub pages.
Really, this does not matter too much as publishing a Dart application to GitHub pages is something of a rarity (I would think) and I have a 14 line Bash script that takes care of it for me.
Day #823
Yeah :/
Thanks for the mention in your write up!
I was using the technique to pull from git as one user and run a dart service process as another user without giving access to the pulling user's home directory.
Ah, makes sense. Like I said in the post, it's not a big deal that it doesn't work for me. I have a 10 line Bash script that does the job. It would be cool if dart pub “just worked” for my GitHub pages use-case, but I can muddle through as-is.
Either way, thanks for the suggestion. I hadn't thought of that and, even though it didn't work in this case, it will come in handy elsewhere. | https://japhr.blogspot.com/2013/07/another-dart-pub-deploy-attempt.html | CC-MAIN-2017-51 | refinedweb | 1,226 | 52.87 |
I.
Join the conversationAdd Comment
hi,
It’s not clear to me if the size presented for the .net 4.0 rtm installer is for the offline version? Because if isn’t, then, well the size is almost equal, in fact the .net 4.0 installer is larger… or am I missing something?
Hi Alex,
The size of .Net Framework 4 in the above chart and table are for offline installers.
The .Net Framework 4 Web Installer is approx 1 MB in size but requires an active internet connection to download the actual installation bits.
Thanks,
Navneet
I build Excel addins for users of Office XP (2002), 2003 and 2007. I have been using a third party tool (ADX) to target them all with one project. Does the Embed Interop Types feature allow me to build a version-neutral addin with just VS2010, or will I still need the third party help?
Hi Bob,
Looks like you are not using VSTO to develop the AddIns and using AddIn Express. I don't know about support for .Net Framework 4 in Add In Express tools you need to talk to them to get more information.
I can talk about VSTO though.
A VSTO AddIn developed for older versions of Office should run as is in later versions of Office without any changes but it won't be able to use the newer features.
The following two blog entries talk more about targeting multiple version of Office using VSTO
blogs.msdn.com/…/targeting-multiple-versions-of-office-without-pias.aspx
blogs.msdn.com/…/can-you-build-one-add-in-for-multiple-versions-of-office.aspx
With VS 2010 you can not develop Office AddIns for Office 2003 or earlier using VSTO, you can use Shared AddIns and COM Shim Wizard though.
By Embedding Interop Types you are no more dependent upon the PIA redistributable and if you are only using the PIA types that are not specific/new to a given version you are good.
Thanks,
Navneet
Thank you for the quick reply.
Most of my customers are still using Office XP (they're a very price sensitive bunch), so VSTO is out. It sounds like I still need the third party solution.
Essentially you have two options here, you can develop Shared AddIns in Visual Studio 2010, or go with ADX.
To learn more about Shared AddIns you may follow the posts below
blogs.msdn.com/…/the-case-for-shared-add-ins.aspx
msdn.microsoft.com/…/aa159894(office.11).aspx
blogs.msdn.com/…/com-shim-wizards-for-vs-2010.aspx
Thanks,
Navneet
Hi!
I wonder how I can upgrade designer-defined files for Excel VSTO such as my ThisWorkbook.designer.cs? I'm asking because after I switched to 4.0 I started getting this error message:
ThisWorkbook.Designer.cs(21,116): error CS0234: The type or namespace name 'Office' does not exist in the namespace 'Microsoft.VisualStudio.Tools' (are you missing an assembly reference?)
If to look at this .cs file we may see there:
[Microsoft.VisualStudio.Tools.Applications.Runtime.StartupObjectAttribute(0)]
[global::System.Security.Permissions.PermissionSetAttribute(global::System.Security.Permissions.SecurityAction.Demand, Name="FullTrust")]
public sealed partial class ThisWorkbook : Microsoft.Office.Tools.Excel.Workbook, Microsoft.VisualStudio.Tools.Office.IOfficeEntryPoint {
What's the best way to change the underlying code?
Thx
Can you share more details about the customization e.g. original target framework version, are you using Visual Studio 2010 RTM?
There are some manual code changes required after retargeting a VSTO customization to .Net Framework 4. Please see the following blog for details
blogs.msdn.com/…/fixing-compile-and-run-time-errors-after-retargeting-vsto-projects-to-the-net-framework-4-mclean-schofield.aspx
Senglory – Please see my response in your thread about this issue on the VSTO forums: social.msdn.microsoft.com/…/fc39c704-1ef2-4810-8e61-a52fe07161a0
It appears I have both the 3.5 SP1 and the 4 client profile. Is it possible to delete the SP1 of 3.5 without ruining what I have in my computer? It is taking a lot of space. I appreciate your enlightenment on this one. Advanced gratitude!
Can I remove 3.5 if I have 4 installed
Visual Studio 2010 does not install .Net Framework 3.5, you can uninstall .Net Framework 3.5 from your developer machine, please take care that you do not have any other apps that require .Net Framework 3.5
Thanks,
Navneet
SO CAN I HAVE NET FRAMWORK 3. AND 4.0 RUNNING ATTHE SAME TIME IN MY COMPUTER?
Plz tell me these difference in briefly
@ JP … yes you can have both .Net Framework 3.5 and 4 installed at the same time.
@ Mohit I did not get your question, please explain.? | https://blogs.msdn.microsoft.com/vsto/2010/04/23/why-should-i-upgrade-from-net-framework-3-5-to-net-framework-4-navneet-gupta/ | CC-MAIN-2016-36 | refinedweb | 788 | 60.01 |
Abstract: Java 1.4 changed inner classes slighty to set the link to the outer object prior to calling super(). This avoids some hard to explain NullPointerExceptions.
First of all, I wish you all the best for the year 2003!
Two of our readers, Dimitris Andreou from Greece and another by
the name of Snow, pointed out that in JDK 1.4 the
NullPointerException has been fixed. Have a look at
Sun's
website. Try compile the following class
with the
-target 1.4 option:
public class NestedBug3 { private Integer wings = new Integer(2); public NestedBug3() { new ComplexBug(); } private class ComplexBug extends Insect { ComplexBug() { System.out.println("Inside ComplexBug Constructor"); } public void printDetails() { System.out.println(wings); } } public static void main(String[] arguments) { new NestedBug3(); } }
When you compile this with
-target 1.4 you will only
be able to run it with a JVM 1.4 and later. Here is the output:
Inside Insect() Constructor 2 Inside ComplexBug Constructor
What happens in the inner class? Let's have a quick look:
jad -noinner NestedBug3$ComplexBug.class class NestedBug3$ComplexBug extends Insect { NestedBug3$ComplexBug(NestedBug3 nestedbug3) { this$0 = nestedbug3; super(); System.out.println("Inside ComplexBug Constructor"); } public void printDetails() { System.out.println(NestedBug3.access$000(this$0)); } private final NestedBug3 this$0; /* synthetic field */ }
You can seen that the assignment of
this$0 and the
call to
super() have been swapped around.
In case you were wondering where to get hold of JAD, I wrote to the author and he told me to tell you to look at JAD. I want to hereby publicly thank Pavel Kuznetsov for making this great tool freely available. I have learnt more from this tool than from any other Java tool.
In case you were wondering why I am spending my New Year's Party writing a newsletter, I just want to get this done quickly so that I can put the follow-up into the 2002 folder ;-)... | https://www.javaspecialists.eu/archive/Issue062b-Follow-Up-and-Happy-New-Year.html | CC-MAIN-2020-45 | refinedweb | 323 | 66.64 |
Sebastian Manczak3,551 Points
Inheritance objective
I'm wondering do I need to create a new class Square and delete the Polygon or add a subclass Square to the existing Polygon? Could I please ask for a couple of hints?
namespace Treehouse.CodeChallenges { class Square { public readonly int SideLength; public Square(int 4) { Sidelength = sidelength; } } }
1 Answer
Steven Parker177,509 Points
The instructions ask you to "Create a new type of polygon called Square." That lets you know that you need to keep Polygon and use it as the base class for Square.
And you'll need to include "sidelength" as the parameter for the constructor. The number 4 will be passed to the "base" constructor to provide the value for "NumSides". | https://teamtreehouse.com/community/inheritance-objective | CC-MAIN-2019-51 | refinedweb | 122 | 68.81 |
Q. Java program to find whether a number is power of two or not.
Here you will find an algorithm and program in Java
Java Program to Find Whether a Given Number is Power of 2
class LFC { static boolean find_power_of_two(int num) { if(num==0) return false; return (int)(Math.ceil((Math.log(num) / Math.log(2)))) == (int)(Math.floor(((Math.log(num) / Math.log(2))))); } public static void main(String[] args) { int num=256; if(find_power_of_two(num)) System.out.printf("%d is the power of two\n",num); else System.out.printf("%d is not the power of two\n",num); } }
Output
256 is the power of two. | https://letsfindcourse.com/java-coding-questions/java-program-to-find-whether-a-no-is-power-of-two | CC-MAIN-2022-27 | refinedweb | 110 | 55.44 |
Why would I use a Carriage return vs a next-line escape sequence?
I guess they have their own uses…
[solved] \n vs \r What's the grammar rule here
Why would I use a Carriage return vs a next-line escape sequence?
I always use \n - you can split strings on \n but apparently not on \r
let input = `this and this and this` console.log(input.split('\n')); console.log(input.split('\r')); console.log(input.split('\u000A')); console.log(input.split('\u000D'));
Afaik, these escapes were once actual control codes for physical line-printers and you would need to send both the ‘line feed’ and the ‘carriage return’ to put the print head at the start of the next line.
Windows uses
\r\n, and you don’t generally use
\r for much except that, specifically for compatibility with Windows.
In turn, it’s like that for compatibility with older tech. Typewriters do a carriage return to return the head to the start of the line, then move the paper up so the head sits on a new line -
\r is carriage return,
\n is new line. Older printers were just electronic typewriters, worked in the same way, MS-DOS supported them.
\n is what you’d use. Any text editor should be able to convert automatically, and web browsers are fine with it (Linux rules the web, not Windows). If you’re building Windows software on non-Windows systems the issue comes into play, or if you’re importing text files created by some windows programs onto a different environment (say a Mac), or vice versa.
You guys are talking about Dot Matrix printers?
As man I remember printing reports for school!
I got it. Thanks guys.
Usually, yes. Occasionally,
\r\n is still useful. For example, if you’re creating a CSV file with JavaScript, the CSV spec requires it:
Each record is located on a separate line, delimited by a line break (CRLF).
CRLF = carriage return (
\r) + line feed (
\n). | https://www.freecodecamp.org/forum/t/solved-n-vs-r-whats-the-grammar-rule-here/162228 | CC-MAIN-2019-13 | refinedweb | 336 | 61.67 |
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