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  • how to read input with multiple lines in java

    - by Gandalf StormCrow
    Hi all, Our professor is making us do some basic programming with java, he gaves a website and everything to register and submit our questions, for today I need to do this one example I feel like I'm on the right track but I just can't figure out the rest .. here is the actualy question : **Sample Input:** 10 12 10 14 100 200 **Sample Output:** 2 4 100 And here is what I've got so far : public class Practice { public static int calculateAnswer(String a, String b) { return (Integer.parseInt(b) - Integer.parseInt(a)); } public static void main(String[] args) { System.out.println(calculateAnswer(args[0], args[1])); } } Now I always get the answer 2 because I'm reading the single line, how can I take all lines into account? thank you For some strange reason everytime I want to execute I get this error: C:\sonic>java Practice.class 10 12 Exception in thread "main" java.lang.NoClassDefFoundError: Fact Caused by: java.lang.ClassNotFoundException: Fact.class at java.net.URLClassLoader$1.run(URLClassLoader.java:20 at java.security.AccessController.doPrivileged(Native M at java.net.URLClassLoader.findClass(URLClassLoader.jav at java.lang.ClassLoader.loadClass(ClassLoader.java:307 at sun.misc.Launcher$AppClassLoader.loadClass(Launcher. at java.lang.ClassLoader.loadClass(ClassLoader.java:248 Could not find the main class: Practice.class. Program will exit. Whosever version of answer I use I get this error, what do I do ? However if I run it in eclipse Run as Run Configuration - Program arguments 10 12 10 14 100 200 I get no output EDIT I have made some progress, at first I was getting the compilation error, then runtime error and now I get wrong answer , so can anybody help me what is wrong with this : import java.io.BufferedReader; import java.io.IOException; import java.io.InputStreamReader; import java.math.BigInteger; public class Practice { public static BigInteger calculateAnswer(String a, String b) { BigInteger ab = new BigInteger(a); BigInteger bc = new BigInteger(b); return bc.subtract(ab); } public static void main(String[] args) throws IOException { BufferedReader stdin = new BufferedReader(new InputStreamReader(System.in)); String line; while ((line = stdin.readLine()) != null && line.length()!= 0) { String[] input = line.split(" "); if (input.length == 2) { System.out.println(calculateAnswer(input[0], input[1])); } } } }

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  • Asynchronous readback from opengl front buffer using multiple PBO's

    - by KillianDS
    I am developing an application that needs to read back the whole frame from the front buffer of an openGL application. I can hijack the application's opengl library and insert my code on swapbuffers. At the moment I am successfully using a simple but excruciating slow glReadPixels command without PBO's. Now I read about using multiple PBO's to speed things up. While I think I've found enough resources to actually program that (isn't that hard), I have some operational questions left. I would do something like this: create a series (e.g. 3) of PBO's use glReadPixels in my swapBuffers override to read data from front buffer to a PBO (should be fast and non-blocking, right?) Create a seperate thread to call glMapBufferARB, once per PBO after a glReadPixels, because this will block until the pixels are in client memory. Process the data from step 3. Now my main concern is of course in steps 2 and 3. I read about glReadPixels used on PBO's being non-blocking, will this be an issue if I issue new opengl commands after that very fast? Will those opengl commands block? Or will they continue (my guess), and if so, I guess only swapbuffers can be a problem, will this one stall or will glReadPixels from front buffer be many times faster than swapping (about each 15-30ms) or, worst case scenario, will swapbuffers be executed while glReadPixels is still reading data to the PBO? My current guess is this logic will do something like this: copy FRONT_BUFFER - generic place in VRAM, copy VRAM-RAM. But I have no idea which of those 2 is the real bottleneck and more, what the influence on the normal opengl command stream is. Then in step 3. Is it wise to do this asynchronously in a thread separated from normal opengl logic? At the moment I think not, It seems you have to restore buffer operations to normal after doing this and I can't install synchronization objects in the original code to temporarily block those. So I think my best option is to define a certain swapbuffer delay before reading them out, so e.g. calling glReadPixels on PBO i%3 and glMapBufferARB on PBO (i+2)%3 in the same thread, resulting in a delay of 2 frames. Also, when I call glMapBufferARB to use data in client memory, will this be the bottleneck or will glReadPixels (asynchronously) be the bottleneck? And finally, if you have some better ideas to speed up frame readback from GPU in opengl, please tell me, because this is a painful bottleneck in my current system. I hope my question is clear enough, I know the answer will probably also be somewhere on the internet but I mostly came up with results that used PBO's to keep buffers in video memory and do processing there. I really need to read back the front buffer to RAM and I do not find any clear explanations about performance in that case (which I need, I cannot rely on "it's faster", I need to explain why it's faster). Thank you

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  • C++ vs Matlab vs Python as a main language for Computer Vision Research

    - by Hough
    Hi all, Firstly, sorry for a somewhat long question but I think that many people are in the same situation as me and hopefully they can also gain some benefit from this. I'll be starting my PhD very soon which involves the fields of computer vision, pattern recognition and machine learning. Currently, I'm using opencv (2.1) C++ interface and I especially like its powerful Mat class and the overloaded operations available for matrix and image operations and seamless transformations. I've also tried (and implemented many small vision projects) using opencv python interface (new bindings; opencv 2.1) and I really enjoy python's ability to integrate opencv, numpy, scipy and matplotlib. But recently, I went back to opencv C++ interface because I felt that the official python new bindings were not stable enough and no overloaded operations are available for matrices and images, not to mention the lack of machine learning modules and slow speeds in certain operations. I've also used Matlab extensively in the past and although I've used mex files and other means to speed up the program, I just felt that Matlab's performance was inadequate for real-time vision tasks, be it for fast prototyping or not. When the project becomes larger and larger, many tasks have to be re-written in C and compiled into Mex files increasingly and Matlab becomes nothing more than a glue language. Here comes the sub-questions: For carrying out research in these fields (machine learning, vision, pattern recognition), what is your main or ideal programming language for rapid prototyping of ideas and testing algorithms contained in papers? For computer vision research work, can you list down the pros and cons of using the following languages? C++ (with opencv + gsl + svmlib + other libraries) vs Matlab (with all its toolboxes) vs python (with the imcomplete opencv bindings + numpy + scipy + matplotlib). Are there computer vision PhD/postgrad students here who are using only C++ (with all its availabe libraries including opencv) without even needing to resort to Matlab or python? In other words, given the current existing computer vision or machine learning libraries, is C++ alone sufficient for fast prototyping of ideas? If you're currently using Java or C# for your research, can you list down the reasons why they should be used and how they compare to other languages in terms of available libraries? What is the de facto vision/machine learning programming language and its associated libraries used in your research group? Thanks in advance. Edit: As suggested, I've opened the question to both academic and non-academic computer vision/machine learning/pattern recognition researchers and groups.

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  • jQuery: how can I clear content without getting the dreaded "stop running this script?" dialog?

    - by Cheeso
    I have a div, that holds a div. like this: <div id='reportHolder' class='column'> <div id='report'> </div> </div> Within the inner div, I add a bunch (7-12) of pairs of a and div elements, like this: <h4><a>Heading1</a></h4> <div> ...content here....</div> The total size of the content, is maybe 200k. Each div just contains a fragment of HTML. After I add all the content, I then create an accordion. like this: $('#report').accordion({collapsible:true, active:false}); This all works fine. The problem is, when I try to clear or remove the report div, it takes a looooooong time, and I get 3 or 4 popups asking "Do you want to stop running this script?" I have tried several ways: option 1: $('#report').accordion('destroy'); $('#report').remove(); $("#reportHolder").html("<div id='report'> </div>"); option 2: $('#report').accordion('destroy'); $('#report').html(''); $("#reportHolder").html("<div id='report'> </div>"); option 3: $('#report').accordion('destroy'); $("#reportHolder").html("<div id='report'> </div>"); No matter what, it hangs for a long while. The call to accordion('destroy') seems to not be the source of the delay. It's the erasure of the html content within the report div. EDIT - fixed typo. ps: this happens on FF3.5 as well as IE8 . Questions: What is taking so long? How can I remove content more quickly?

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  • Run bat file in Java and wait 2

    - by Savvas Dalkitsis
    This is a followup question to my other question : http://stackoverflow.com/questions/2434125/run-bat-file-in-java-and-wait The reason i am posting this as a separate question is that the one i already asked was answered correctly. From some research i did my problem is unique to my case so i decided to create a new question. Please go read that question before continuing with this one as they are closely related. Running the proposed code blocks the program at the waitFor invocation. After some research i found that the waitFor method blocks if your process has output that needs to be proccessed so you should first empty the output stream and the error stream. I did those things but my method still blocks. I then found a suggestion to simply loop while waiting the exitValue method to return the exit value of the process and handle the exception thrown if it is not, pausing for a brief moment as well so as not to consume all the CPU. I did this: import java.io.BufferedReader; import java.io.IOException; import java.io.InputStreamReader; public class Test { public static void main(String[] args) { try { Process p = Runtime.getRuntime().exec( "cmd /k start SQLScriptsToRun.bat" + " -UuserName -Ppassword" + " projectName"); final BufferedReader input = new BufferedReader(new InputStreamReader(p.getInputStream())); final BufferedReader error = new BufferedReader(new InputStreamReader(p.getErrorStream())); new Thread(new Runnable() { @Override public void run() { try { while (input.readLine()!=null) {} } catch (IOException e) { e.printStackTrace(); } } }).start(); new Thread(new Runnable() { @Override public void run() { try { while (error.readLine()!=null) {} } catch (IOException e) { e.printStackTrace(); } } }).start(); int i = 0; boolean finished = false; while (!finished) { try { i = p.exitValue(); finished = true; } catch (IllegalThreadStateException e) { e.printStackTrace(); try { Thread.sleep(500); } catch (InterruptedException e1) { e1.printStackTrace(); } } } System.out.println(i); } catch (IOException e) { e.printStackTrace(); } } } but my process will not end! I keep getting this error: java.lang.IllegalThreadStateException: process has not exited Any ideas as to why my process will not exit? Or do you have any libraries to suggest that handle executing batch files properly and wait until the execution is finished?

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  • Opera Mobile, offline web app development, and memory

    - by Jake Krohn
    I'm developing a data collection app for use on a HP iPAQ 211. I'm doing it as an offline web app (go with what you know) using Opera Mobile 9.7 and Google Gears. Being it is an offline app, it is very dependent on Javascript for much of its behavior. I'm using the LocalServer, Database, and Geolocation components of Gears, as well as the JQuery core and a couple of plugins for form validation and other usability tweaks (no jQuery UI). I've tried to be conservative with my programming style and free up or close resources whenever possible, but Opera just slowly dies after about 10-20 minutes of use. The Javascript engine stops responding, pages only half-load, and eventually stop loading completely. I'm guessing it's a resource issue. Quitting and relaunching the browser solves the problem, but only temporarily. The iPAQ ships with 128 MB of RAM, about 85-87 MB of which is available immediately after a reset. With only Opera running, there still remains about 50 MB that is left unused. My questions are thus: Is it possible to get Opera to address this unused RAM? Are there configuration settings in Opera or in the Windows Registry itself that will help improve performance? I know where to tweak, but the descriptions of the opera:config variables that I've found are less than helpful. Is is laughable to ask about memory management and jQuery in the same sentence? If not, does anyone have any suggestions? Finally, are my plans too ambitious, given the platform I have to work with? I know that Gears and Windows Mobile 6 are on their way out, but they (theoretically) suffice for what I need to do. I could ditch them in favor of an iPhone/iPod Touch, Mobile Safari, and HTML5 but I'd like to try to make this work first. I didn't think that Opera was a dog when it comes to JS performance, but perhaps it's worse than I thought. That this motley collection of technologies works at all is a minor miracle, but it needs to be faster and more stable. I appreciate any suggestions.

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  • How to implement a SIMPLE "You typed ACB, did you mean ABC?"

    - by marcgg
    I know this is not a straight up question, so if you need me to provide more information about the scope of it, let me know. There are a bunch of questions that address almost the same issue (they are linked here), but never the exact same one with the same kind of scope and objective - at least as far as I know. Context: I have a MP3 file with ID3 tags for artist name and song title. I have two tables Artists and Songs The ID3 tags might be slightly off (e.g. Mikaell Jacksonne) I'm using ASP.NET + C# and a MSSQL database I need to synchronize the MP3s with the database. Meaning: The user launches a script The script browses through all the MP3s The script says "Is 'Mikaell Jacksonne' 'Michael Jackson' YES/NO" The user pick and we start over Examples of what the system could find: In the database... SONGS = {"This is a great song title", "This is a song title"} ARTISTS = {"Michael Jackson"} Outputs... "This is a grt song title" did you mean "This is a great song title" ? "This is song title" did you mean "This is a song title" ? "This si a song title" did you mean "This is a song title" ? "This si song a title" did you mean "This is a song title" ? "Jackson, Michael" did you mean "Michael Jackson" ? "JacksonMichael" did you mean "Michael Jackson" ? "Michael Jacksno" did you mean "Michael Jackson" ? etc. I read some documentation from this /how-do-you-implement-a-did-you-mean and this is not exactly what I need since I don't want to check an entire dictionary. I also can't really use a web service since it's depending a lot on what I already have in my database. If possible I'd also like to avoid dealing with distances and other complicated things. I could use the google api (or something similar) to do this, meaning that the script will try spell checking and test it with the database, but I feel there could be a better solution since my database might end up being really specific with weird songs and artists, making spell checking useless. I could also try something like what has been explained on this post, using Soundex for c#. Using a regular spell checker won't work because I won't be using words but names and 'titles'. So my question is: is there a relatively simple way of doing this, and if so, what is it? Any kind of help would be appreciated. Thanks!

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  • int, short, byte performance in back-to-back for-loops

    - by runrunraygun
    (background: http://stackoverflow.com/questions/1097467/why-should-i-use-int-instead-of-a-byte-or-short-in-c) To satisfy my own curiosity about the pros and cons of using the "appropriate size" integer vs the "optimized" integer i wrote the following code which reinforced what I previously held true about int performance in .Net (and which is explained in the link above) which is that it is optimized for int performance rather than short or byte. DateTime t; long a, b, c; t = DateTime.Now; for (int index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } a = DateTime.Now.Ticks - t.Ticks; t = DateTime.Now; for (short index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } b=DateTime.Now.Ticks - t.Ticks; t = DateTime.Now; for (byte index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } c=DateTime.Now.Ticks - t.Ticks; Console.WriteLine(a.ToString()); Console.WriteLine(b.ToString()); Console.WriteLine(c.ToString()); This gives roughly consistent results in the area of... ~950000 ~2000000 ~1700000 which is in line with what i would expect to see. However when I try repeating the loops for each data type like this... t = DateTime.Now; for (int index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } for (int index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } for (int index = 0; index < 127; index++) { Console.WriteLine(index.ToString()); } a = DateTime.Now.Ticks - t.Ticks; the numbers are more like... ~4500000 ~3100000 ~300000 Which I find puzzling. Can anyone offer an explanation? NOTE: In the interest of compairing like for like i've limited the loops to 127 because of the range of the byte value type. Also this is an act of curiosity not production code micro-optimization.

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  • Calculate year for end date: PostgreSQL

    - by Dave Jarvis
    Background Users can pick dates as shown in the following screen shot: Any starting month/day and ending month/day combinations are valid, such as: Mar 22 to Jun 22 Dec 1 to Feb 28 The second combination is difficult (I call it the "tricky date scenario") because the year for the ending month/day is before the year for the starting month/day. That is to say, for the year 1900 (also shown selected in the screen shot above), the full dates would be: Dec 22, 1900 to Feb 28, 1901 Dec 22, 1901 to Feb 28, 1902 ... Dec 22, 2007 to Feb 28, 2008 Dec 22, 2008 to Feb 28, 2009 Problem Writing a SQL statement that selects values from a table with dates that fall between the start month/day and end month/day, regardless of how the start and end days are selected. In other words, this is a year wrapping problem. Inputs The query receives as parameters: Year1, Year2: The full range of years, independent of month/day combination. Month1, Day1: The starting day within the year to gather data. Month2, Day2: The ending day within the year (or the next year) to gather data. Previous Attempt Consider the following MySQL code (that worked): end_year = start_year + greatest( -1 * sign( datediff( date( concat_ws('-', year, end_month, end_day ) ), date( concat_ws('-', year, start_month, start_day ) ) ) ), 0 ) How it works, with respect to the tricky date scenario: Create two dates in the current year. The first date is Dec 22, 1900 and the second date is Feb 28, 1900. Count the difference, in days, between the two dates. If the result is negative, it means the year for the second date must be incremented by 1. In this case: Add 1 to the current year. Create a new end date: Feb 28, 1901. Check to see if the date range for the data falls between the start and calculated end date. If the result is positive, the dates have been provided in chronological order and nothing special needs to be done. This worked in MySQL because the difference in dates would be positive or negative. In PostgreSQL, the equivalent functionality always returns a positive number, regardless of their relative chronological order. Question How should the following (broken) code be rewritten for PostgreSQL to take into consideration the relative chronological order of the starting and ending month/day pairs (with respect to an annual temporal displacement)? SELECT m.amount FROM measurement m WHERE (extract(MONTH FROM m.taken) >= month1 AND extract(DAY FROM m.taken) >= day1) AND (extract(MONTH FROM m.taken) <= month2 AND extract(DAY FROM m.taken) <= day2) Any thoughts, comments, or questions? (The dates are pre-parsed into MM/DD format in PHP. My preference is for a pure PostgreSQL solution, but I am open to suggestions on what might make the problem simpler using PHP.) Versions PostgreSQL 8.4.4 and PHP 5.2.10

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  • Showing login view controller before main tab bar controller

    - by Padawan
    I'm creating an iPad app with a tab bar controller that requires login. So on launch, I want to show a LoginViewController and if login is successful, then show the tab bar controller. This is how I implemented an initial test version (left out some typical header stuff, etc)... AppDelegate.h: @interface AppDelegate_Pad : NSObject <UIApplicationDelegate, LoginViewControllerDelegate> { UIWindow *window; UITabBarController *tabBarController; } @property (nonatomic, retain) IBOutlet UIWindow *window; @property (nonatomic, retain) IBOutlet UITabBarController *tabBarController; @end AppDelegate.m: @implementation AppDelegate_Pad @synthesize window; @synthesize tabBarController; - (BOOL)application:(UIApplication *)application didFinishLaunchingWithOptions:(NSDictionary *)launchOptions { LoginViewController_Pad *lvc = [[LoginViewController_Pad alloc] initWithNibName:@"LoginViewController_Pad" bundle:nil]; lvc.delegate = self; [window addSubview:lvc.view]; //[lvc release]; [window makeKeyAndVisible]; return YES; } - (void)loginViewControllerDidFinish:(LoginViewController_Pad *)loginViewController { [window addSubview:tabBarController.view]; } - (void)dealloc {...} @end LoginViewController_Pad.h: @protocol LoginViewControllerDelegate; @interface LoginViewController_Pad : UIViewController { id<LoginViewControllerDelegate> delegate; } @property (nonatomic, assign) id <LoginViewControllerDelegate> delegate; - (IBAction)buttonPressed; @end @protocol LoginViewControllerDelegate -(void)loginViewControllerDidFinish:(LoginViewController_Pad *)loginViewController; @end LoginViewController_Pad.m: @implementation LoginViewController_Pad @synthesize delegate; ... - (IBAction)buttonPressed { [self.view removeFromSuperview]; [self.delegate loginViewControllerDidFinish:self]; } ... @end So the app delegate adds the login view controller's view on launch and waits for login to call "did finish" using a delegate. The login view controller calls removeFromSuperView before it calls didFinish. The app delegate then calls addSubView on the tab bar controller's view. If you made it up to this point, thanks, and I have three questions: MAIN QUESTION: Is this the right way to show a view controller before the app's main tab bar controller is displayed? Even though it seems to work, is it a proper way to do it? If I comment out the "lvc release" in the app delegate then the app crashes with EXC_BAD_ACCESS when the button on the login view controller is pressed. Why? With the "lvc release" commented out everything seems to work but on the debugger console it writes this message when the app delegate calls addSubView for the tab bar controller: Using two-stage rotation animation. To use the smoother single-stage animation, this application must remove two-stage method implementations. What does that mean and do I need to worry about it?

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  • Haskell: Left-biased/short-circuiting function

    - by user2967411
    Two classes ago, our professor presented to us a Parser module. Here is the code: module Parser (Parser,parser,runParser,satisfy,char,string,many,many1,(+++)) where import Data.Char import Control.Monad import Control.Monad.State type Parser = StateT String [] runParser :: Parser a -> String -> [(a,String)] runParser = runStateT parser :: (String -> [(a,String)]) -> Parser a parser = StateT satisfy :: (Char -> Bool) -> Parser Char satisfy f = parser $ \s -> case s of [] -> [] a:as -> [(a,as) | f a] char :: Char -> Parser Char char = satisfy . (==) alpha,digit :: Parser Char alpha = satisfy isAlpha digit = satisfy isDigit string :: String -> Parser String string = mapM char infixr 5 +++ (+++) :: Parser a -> Parser a -> Parser a (+++) = mplus many, many1 :: Parser a -> Parser [a] many p = return [] +++ many1 p many1 p = liftM2 (:) p (many p) Today he gave us an assignment to introduce "a left-biased, or short-circuiting version of (+++)", called (<++). His hint was for us to consider the original implementation of (+++). When he first introduced +++ to us, this was the code he wrote, which I am going to call the original implementation: infixr 5 +++ (+++) :: Parser a -> Parser a -> Parser a p +++ q = Parser $ \s -> runParser p s ++ runParser q s I have been having tons of trouble since we were introduced to parsing and so it continues. I have tried/am considering two approaches. 1) Use the "original" implementation, as in p +++ q = Parser $ \s - runParser p s ++ runParser q s 2) Use the final implementation, as in (+++) = mplus Here are my questions: 1) The module will not compile if I use the original implementation. The error: Not in scope: data constructor 'Parser'. It compiles fine using (+++) = mplus. What is wrong with using the original implementation that is avoided by using the final implementation? 2) How do I check if the first Parser returns anything? Is something like (not (isNothing (Parser $ \s - runParser p s) on the right track? It seems like it should be easy but I have no idea. 3) Once I figure out how to check if the first Parser returns anything, if I am to base my code on the final implementation, would it be as easy as this?: -- if p returns something then p <++ q = mplus (Parser $ \s -> runParser p s) mzero -- else (<++) = mplus Best, Jeff

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  • Google maps API - info window height and panning

    - by Tim Fountain
    I'm using the Google maps API (v2) to display a country overlay over a world map. The data comes from a KML file, which contains coords for the polygons along with a HTML description for each country. This description is displayed in the 'info window' speech bubble when that country is clicked on. I had some trouble initially as the info windows were not expanding to the size of the HTML content they contained, so the longer ones would spill over the edges (this seems to be a common problem). I was able to work around this by resetting the info window to a specific height as follows: GEvent.addListener(map, "infowindowopen", function(iw) { iw = map.getInfoWindow(); iw.reset(iw.getPoint(), iw.getTabs(), new GSize(300, 295), null, null); }); Not ideal, but it works. However now, when the info windows are opened the top part of them is sometimes obscured by the edges of the map, as the map does not pan to a position where all of the content can be viewed. So my questions: Is there any way to get the info windows to automatically use a height appropriate to their content, to avoid having to fix to a set pixel height? If fixing the height is the only option, is there any way to get the map to pan to a more appropriate position when the info windows open? I know that the map class has a panTo() method, but I can't see a way to calculate what the correct coords would be. Here's my full init code: google.load("maps", "2.x"); // Call this function when the page has been loaded function initialize() { var map = new google.maps.Map2(document.getElementById("map"), {backgroundColor:'#99b3cc'}); map.addControl(new GSmallZoomControl()); map.setCenter(new google.maps.LatLng(29.01377076013671, -2.7866649627685547), 2); gae_countries = new GGeoXml("http://example.com/countries.kmz"); map.addOverlay(gae_countries); GEvent.addListener(map, "infowindowopen", function(iw) { iw = map.getInfoWindow(); iw.reset(iw.getPoint(), iw.getTabs(), new GSize(300, 295), null, null); }); } google.setOnLoadCallback(initialize);

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  • Does weak typing offer any advantages?

    - by sub
    Don't confuse this with static vs. dynamic typing! You all know JavaScripts/PHPs infamous type systems: PHP example: echo "123abc"+2; // 125 - the reason for this is explained // in the PHP docs but still: This hurts echo "4"+1; // 5 - Oh please echo "ABC"*5; // 0 - WTF // That's too much, seriously now. // This here might be actually a use for weak typing, but no - // it has to output garbage. JavaScript example: // A good old JavaScript, maybe you'll do better? alert("4"+1); // 51 - Oh come on. alert("abc"*3); // NaN - What the... // Have your creators ever heard of the word "consistence"? Python example: # Python's type system is actually a mix # It spits errors on senseless things like the first example below AND # allows intelligent actions like the second example. >>> print("abc"+1) Traceback (most recent call last): File "<pyshell#2>", line 1, in <module> print("abc"+1) TypeError: Can't convert 'int' object to str implicitly >>> print("abc"*5) abcabcabcabcabc Ruby example: puts 4+"1" // Type error - as supposed puts "abc"*4 // abcabcabcabc - makes sense After these examples it should be clear that PHP/JavaScript probably have the most inconsistent type systems out there. This is a fact and really not subjective. Now when having a closer look at the type systems of Ruby and Python it seems like they are having a much more intelligent and consistent type system. I think these examples weren't really necessary as we all know that PHP/JavaScript have a weak and Python/Ruby have a strong type system. I just wanted to mention why I'm asking this. Now I have two questions: When looking at those examples, what are the advantages of PHPs and JavaScripts type systems? I can only find downsides: They are inconsistent and I think we know that this is not good Types conversions are hardly controllable Bugs are more likely to happen and much harder to spot Do you prefer one of the both systems? Why? Personally I have worked with PHP, JavaScript and Python so far and must say that Pythons type system has really only advantages over PHPs and JavaScripts. Does anybody here not think so? Why then?

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  • Best way to handle multiple tables to replace one big table in Rails? (e.g. 'Books1', 'Books2', etc.

    - by mikep
    Hello, I've decided to use multiple tables for an entity (e.g. Books1, Books2, Books3, etc.), instead of just one main table which could end up having a lot of rows (e.g. just Books). I'm doing this to try and to avoid a potential future performance drop that could come with having too many rows in one table. With that, I'm looking for a good way to handle this in Rails, mainly by trying to avoid loading a bunch of unused associations. (I know that I could use a partition for this, but, for now, I've decided to go the 'multiple tables' route.) Each user has their books placed into a specific table. The actual book table is chosen when the user is created, and all of their books go into the same table. I'm going to split the adds across the tables. The goal is to try and keep each table pretty much even -- but that's a different issue. One thing I don't particularly want to have is a bunch of unused associations in the User class. Right now, it looks like I'd have to do the following: class User < ActiveRecord::Base has_many :books1, :books2, :books3, :books4, :books5 end class Books1 < ActiveRecord::Base belongs_to :user end class Books2 < ActiveRecord::Base belongs_to :user end class Books3 < ActiveRecord::Base belongs_to :user end I'm assuming that the main performance hit would come in terms of memory and possibly some method call overhead for each User object, since it has to load all of those associations, which in turn creates all of those nice, dynamic model accessor methods like User.find_by_. But for each specific user, only one of the book tables would be usable/applicable, since all of a user's books are stored in the same table. So, only one of the associations would be in use at any time and any other has_many :bookX association that was loaded would be a waste. For example, with a user.id of 2, I'd only need books3.find_by_author('Author'), but the way I'm thinking of setting this up, I'd still have access to Books1..n. I don't really know Ruby/Rails does internally with all of those has_many associations though, so maybe it's not so bad. But right now I'm thinking that it's really wasteful, and that there may just be a better, more efficient way of doing this. So, a few questions: 1) Is there's some sort of special Ruby/Rails methodology that could be applied to this 'multiple tables to represent one entity' scheme? Are there any 'best practices' for this? 2) Is it really bad to have so many unused has_many associations for each object? Is there a better way to do this? 3) Does anyone have any advice on how to abstract the fact that there's multiple book tables behind a single books model/class? For example, so I can call books.find_by_author('Author') instead of books3.find_by_author('Author'). Thank you!

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  • AutoScaleMode problems with changed default font

    - by Doc Brown
    Hi, I have some problems with the Form.AutoScaleMode property together with fixed size controls, when using a non-default font. I boiled it down to a simple test application (WinForms 2.0) with only one form, some fixed size controls and the following properties: class Form1 : Form { // ... private void InitializeComponent() { // ... this.AutoScaleDimensions = new System.Drawing.SizeF(96F, 96F); this.AutoScaleMode = System.Windows.Forms.AutoScaleMode.Dpi; this.Font = new System.Drawing.Font("Tahoma", 9.25F); // ... } } Under 96dpi, Windows XP, the form looks correctly like this 96 dpi example. Under 120 dpi, Windows XP, the the Windows Forms autoscaling feature produces this 120 dpi example. As you can see, groupboxes, buttons, list or tree views are scaled correctly, multiline text boxes get too big in the vertical axis, and a fixed size label does not scale correctly in both vertical and horizontal direction. Seems to be bug in the .NET framework? Using the default font (Microsoft Sans Serif 8.25pt), this problem does not occur. Using AutoScaleMode=Font (with adequate AutoScaleDimensions, of course) either does not scale at all or scales exactly like seen above, depending on when the Font is set (before or after the change of AutoScaleMode). The problem is not specific to the "Tahoma" Font, it occurs also with Microsoft Sans Serif, 9.25pt. And yes, i already read this SO post http://stackoverflow.com/questions/2114857/high-dpi-problems but it does not really help me. Any suggestions how to come around this? EDIT: I changed my image hoster, hope this one works better. EDIT2: Some additional information about my intention: I have about 50 already working fixed size dialogs with several hundreds of properly placed, fixed size controls. They were migrated from an older C++ GUI framework to C#/Winforms, that's why they are all fixed-size. All of them look fine with 96 dpi using a 9.25pt font. Under the old framework, scaling to 120 dpi worked fine - all fixed size controls scaled equal in both dimensions. Last week, we detected this strange scaling behaviour under WinForms when switching to 120 dpi. You can imagine that most of our dialogs now look very bad under 120 dpi. We are looking for a solution that avoids a complete redesign all those dialogs.

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  • SmoApplication.EnumAvailableSqlServers returns server names but not instance names (but only on one

    - by Matma
    Hi, There are a number of questions about this and a number of possible causes and thus far ive tried them all with no success. situation: i have an app that needs a db to work, onstartup it does a SmoApplication.EnumAvailableSqlServers(false) to get all the instances on the network, shows the user a dropdown, they pick one and i go connect to my db on that server. all good problem: this works on my machine, the guys next to me and others. HOWEVER it doesnt work on one of the tech guys machines (and potentially others). we are all on the same network domain, physically connected (no wireless), all logged on with network user names, all running the same sql express 2005 sp3, though im using win7 the other guys are running xppro. MSSMS on all machines can see all the instances when you select "Browse for more". yet on this one tech guys machine it lists his local instance (since its hardcoded to) and all the network servers, but has no instances names? i.e. .sqlexpress server1 server2 server3 server4 but on my machine and others we get: .sqlexpress server1/sqlexpress server2/sqlexpress server3/sqlexpress server4/sqlexpress the code im using: ' .... some code ' this populates my datatable dtServers = SmoApplication.EnumAvailableSqlServers(False) '.... some code '.... then later i ShowServers(...) Private dtServers As DataTable = Nothing Private Sub ShowServers(ByVal SQLInstance As String) ' Create a DataTable where we enumerate the available servers cmbServer.Items.Clear() cmbDatabase.Items.Clear() ' If there are any (network listed) servers at all If (dtServers.Rows.Count > 0) Then ' Loop through each server in the DataTable For Each drServer As DataRow In dtServers.Rows ' Add the name to the combobox cmbServer.Items.Add(drServer("Server") & "\" & drServer("Instance")) Next End If 'To make life simpler (add the local instance of sql express): cmbServer.Items.Add(SQLInstance) ' select first item If cmbServer.Items.Count > 0 Then cmbServer.SelectedIndex = 0 End If End Sub now i know this uses udp and its not 100%, but how come his machine is 100% consistent in not showing remote instances, and mine is 100 consistent showing them. even a udl file on his desktop cant see them, regarldess of provider i choose to use? some of the suggestions are to uninstall and re-install, but that doesnt seem like a solution as i (and most others) can see the instances, but one guy cant. this suggests its not the remote sql server but rather the local machine. Notes: ive tried firewall 1433, 1434 i can connect using a udl with full SERVERNAME\INSTANCENAME the browser service is running locally and on the remote machine ive tried stopping and restarting both the browser service on the local and remote machine. Ideas?

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  • MooseX::Types declaration issue, tight test case :)

    - by TJ Thompson
    So after an embarrassing amount of time debugging, I've finally stripped this issue ([http://stackoverflow.com/questions/4621589/perl-moose-typedecorator-error-how-do-i-debug][1]) down to a simple test case. I would humbly request some help understanding why it's failing :) Here is the error message I'm getting: plxc16479 $h2/tmp/tmp18.pl This method [new] requires a single argument. at /nfs/pdx/disks/nehalem.pde.077/perl/5.12.2/lib64/site_perl/MooseX/Types/TypeDecorator.pm line 91 MooseX::Types::TypeDecorator::new('MooseX::Types::TypeDecorator=HASH(0x655b90)') called at /nfs/pdx/disks/nehalem.pde.077/projects/lib/Program-Plist-Pl/lib/Program/Plist/Pl.pm line 10 Program::Plist::Pl::BUILD('Program::Plist::Pl=HASH(0x63d478)', 'HASH(0x63d220)') called at generated method (unknown origin) line 29 Program::Plist::Pl::new('Program::Plist::Pl') called at /nfs/pdx/disks/nehalem.pde.077/tmp/tmp18.pl line 10 Wrapper test script: use strict; use warnings; BEGIN {push(@INC, split(':', $ENV{PERL_TEST_LIBS}))}; use Program::Plist::Pl; my $obj = Program::Plist::Pl->new(); Program::Plist::Pl file: package Program::Plist::Pl; use Moose; use namespace::autoclean; use Program::Types qw(Pattern); # <-- Removing this fixes error use Program::Plist::Pl::Pattern; sub BUILD { my $pattern_obj = Program::Plist::Pl::Pattern->new(); } __PACKAGE__->meta->make_immutable; 1; Program::Types file: package Program::Types; use MooseX::Types -declare => [qw(Pattern)]; class_type Pattern, {class => 'Program::Plist::Pl::Pattern'}; 1; And the Program::Plist::Pl::Pattern file: package Program::Plist::Pl::Pattern; use Moose; use namespace::autoclean; __PACKAGE__->meta->make_immutable; 1; Notes: While I don't need the Pattern type from Program::Types in the above code, I do in other code that is stripped out. The PERL_TEST_LIBS env var I'm pulling INC paths from only contains paths to the project modules. There are no other modules loaded from these paths. It appears the MooseX::Types definition for Pattern is causing problems, but I'm not sure why. Documentation shows the syntax I am using, but it's possible I'm misusing class_type as there isn't much said about it. Intent is to be able to use Pattern for type checking via MooseX::Params::Validate to verify the argument is a 'Program::Plist::Pl::Program' object. I've found that removing the intervening class Program::Plist::Pl from the equation by directly calling Pattern-new from the tmp18.pl wrapper results in no error, even when the Program::Types Pattern type is imported.

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  • Please Explain Drupal schema and drupal_write_record

    - by Aaron
    Hi. A few questions. 1) Where is the best place to populate a new database table when a module is first installed, enabled? I need to go and get some data from an external source and want to do it transparently when the user installs/enables my custom module. I create the schema in {mymodule}_schema(), do drupal_install_schema({tablename}); in hook_install. Then I try to populate the table in hook_enable using drupal_write_record. I confirmed the table was created, I get no errors when hook_enable executes, but when I query the new table, I get no rows back--it's empty. Here's one variation of the code I've tried: /** * Implementation of hook_schema() */ function ncbi_subsites_schema() { // we know it's MYSQL, so no need to check $schema['ncbi_subsites_sites'] = array( 'description' => 'The base table for subsites', 'fields' => array( 'site_id' => array( 'description' => 'Primary id for site', 'type' => 'serial', 'unsigned' => TRUE, 'not null' => TRUE, ), // end site_id 'title' => array( 'description' => 'The title of the subsite', 'type' => 'varchar', 'length' => 255, 'not null' => TRUE, 'default' => '', ), //end title field 'url' => array( 'description' => 'The URL of the subsite in Production', 'type' => 'varchar', 'length' => 255, 'default' => '', ), //end url field ), //end fields 'unique keys' => array( 'site_id'=> array('site_id'), 'title' => array('title'), ), //end unique keys 'primary_key' => array('site_id'), ); // end schema return $schema; } Here's hook_install: function ncbi_subsites_install() { drupal_install_schema('ncbi_subsites'); } Here's hook_enable: function ncbi_subsites_enable() { drupal_get_schema('ncbi_subsites_site'); // my helper function to get data for table (not shown) $subsites = ncbi_subsites_get_subsites(); foreach( $subsites as $name=>$attrs ) { $record = new stdClass(); $record->title = $name; $record->url = $attrs['homepage']; drupal_write_record( 'ncbi_subsites_sites', $record ); } } Can someone tell me what I'm missing?

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  • BindException with INTERNET permission requested

    - by Mondain
    I have seen several questions regarding SocketException when using Android, but none of them cover the BindException that I get even with the INTERNET permission specified in my manifest. Here is part of my manifest: <uses-permission android:name="android.permission.INTERNET"></uses-permission> <uses-permission android:name="android.permission.ACCESS_NETWORK_STATE"></uses-permission> <uses-permission android:name="android.permission.ACCESS_WIFI_STATE"></uses-permission> <uses-permission android:name="android.permission.READ_OWNER_DATA"></uses-permission> <uses-permission android:name="android.permission.READ_PHONE_STATE"></uses-permission> <uses-permission android:name="android.permission.ACCOUNT_MANAGER"></uses-permission> <uses-permission android:name="android.permission.AUTHENTICATE_ACCOUNTS"></uses-permission> Here is the relevant portion of my LogCat output: 04-22 14:49:06.117: DEBUG/MyLibrary(4844): Address to bind: 192.168.1.14 port: 843 04-22 14:49:06.197: WARN/System.err(4844): java.net.BindException: Permission denied (maybe missing INTERNET permission) 04-22 14:49:06.207: WARN/System.err(4844): at org.apache.harmony.luni.platform.OSNetworkSystem.socketBindImpl(Native Method) 04-22 14:49:06.207: WARN/System.err(4844): at org.apache.harmony.luni.platform.OSNetworkSystem.bind(OSNetworkSystem.java:107) 04-22 14:49:06.217: WARN/System.err(4844): at org.apache.harmony.luni.net.PlainSocketImpl.bind(PlainSocketImpl.java:184) 04-22 14:49:06.217: WARN/System.err(4844): at java.net.ServerSocket.bind(ServerSocket.java:414) 04-22 14:49:06.227: WARN/System.err(4844): at org.apache.harmony.nio.internal.ServerSocketChannelImpl$ServerSocketAdapter.bind(ServerSocketChannelImpl.java:213) 04-22 14:49:06.227: WARN/System.err(4844): at java.net.ServerSocket.bind(ServerSocket.java:367) 04-22 14:49:06.237: WARN/System.err(4844): at org.apache.harmony.nio.internal.ServerSocketChannelImpl$ServerSocketAdapter.bind(ServerSocketChannelImpl.java:283) 04-22 14:49:06.237: WARN/System.err(4844): at mylibrary.net.PolicyConnection$PolicyServerWorker.(PolicyConnection.java:201) I Really hope this is a simple problem and not something complicated by the fact that the binding is occurring within a worker thread on a port less than 1024. Update Looks as if this is a privileged port issue, anyone know how to bind to ports lower than 1024 in Android? SelectorProvider provider = SelectorProvider.provider(); try { ServerSocketChannel channel = provider.openServerSocketChannel(); policySocket = channel.socket(); Log.d("MyLibrary", "Address to bind: " + device.getAddress().getAddress() + " port: 843"); InetSocketAddress addr = new InetSocketAddress(InetAddress.getByName(device.getAddress().getAddress()), 843); policySocket.bind(addr); policySocket.setReuseAddress(true); policySocket.setReceiveBufferSize(256); } catch (Exception e) { e.printStackTrace(); }

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  • Where would I implement this array to pass?

    - by Keeano Martin
    I currently build an NSMutableArray in Class A.m within the ViewDidLoad Method. - (void)viewDidLoad { [super viewDidLoad]; //Question Array Setup and Alloc stratToolsDict = [[NSMutableDictionary alloc] initWithObjectsAndKeys:countButton,@"count",camerButton,@"camera",videoButton,@"video",textButton,@"text",probeButton,@"probe", nil]; stratTools = [[NSMutableArray alloc] initWithObjects:@"Tools",stratToolsDict, nil]; stratObjectsDict = [[NSMutableDictionary alloc]initWithObjectsAndKeys:stratTools,@"Strat1",stratTools,@"Strat2",stratTools,@"Strat3",stratTools,@"Strat4", nil]; stratObjects = [[NSMutableArray alloc]initWithObjects:@"Strategies:",stratObjectsDict,nil]; QuestionDict = [[NSMutableDictionary alloc]initWithObjectsAndKeys:stratObjects,@"Question 1?",stratObjects,@"Question 2?",stratObjects,@"Question 3?",stratObjects,@"Question 4?",stratObjects,@"Question 5?", nil]; //add strategys to questions QuestionsList = [[NSMutableArray alloc]init]; for (int i = 0; i < 1; i++) { [QuestionsList addObject:QuestionDict]; } NSLog(@"Object: %@",QuestionsList); At the end of this method you will see QuestionsList being initialized and now I need to send this Array to Class B. So I place its setters and getters using the @property and @Synthesize method. Class A.h @property (retain, nonatomic) NSMutableDictionary *stratToolsDict; @property (retain, nonatomic) NSMutableArray *stratTools; @property (retain, nonatomic) NSMutableArray *stratObjects; @property (retain, nonatomic) NSMutableDictionary *QuestionDict; @property (retain, nonatomic) NSMutableArray *QuestionsList; Class A.m @synthesize QuestionDict; @synthesize stratToolsDict; @synthesize stratObjects; @synthesize stratTools; @synthesize QuestionsList; I use the property method because I am going to call this variable from Class B and want to be able to assign it to another NSMutableArray. I then add the @property and @class for Class A to Class B.h as well as declare the NSMutableArray in the @interface. #import "Class A.h" @class Class A; @interface Class B : UITableViewController<UITableViewDataSource, UITableViewDelegate>{ NSMutableArray *QuestionList; Class A *arrayQuestions; } @property Class A *arrayQuestions; Then I call NSMutableArray from Class A in the Class B.m -(id)initWithStyle:(UITableViewStyle)style { if ([super initWithStyle:style] != nil) { //Make array arrayQuestions = [[Class A alloc]init]; QuestionList = arrayQuestions.QuestionsList; Right after this I Log the NSMutableArray to view values and check that they are there and it returns NIL. //Log test NSLog(@"QuestionList init method: %@",QuestionList); Info about Class B- Class B is a UIPopOverController for Class A, Class B has one View which holds a UITableView which I have to populate the results of Class A's NSMutableArray. Why is the NsMutableArray coming back as NIL? Ultimately would like some help figuring it out as well, it seems to really have me confused. Help is greatly appreciated!!

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  • Asp.net MVC and MOSS 2010 integration

    - by Robert Koritnik
    Just a sidenote: I'm not sure whether I should post this to serverfault as well, because some MOSS admin may have some info for me as well? A bit of explanation first (without Asp.net MVC) Is it possible to integrate the two? Is it possible to write an application that would share at least credential information with MOSS? I have to write a MOSS application that has to do with these technologies: MOSS 2010 Personal client certificates authentication (most probably on USB keys) Active Directory Federation Services Separate SQL DB that would serve application specific data (separate as not being part of MOSS DB) How should it work? Users should authenticate using personal certificates into MOSS 2010 There would be a certain part of MOSS that would be related to my custom application This application should only authorize certain users via AD FS - I guess these users should have a certain security claim attached to them This application should manage users (that have access to this app) with additional (app specific) security claims related to this application (as additional application level authorization rights for individual application parts) This application should use custom SQL 2008 DB heavily with its own data This application should have the possibility to integrate with external systems as well (Exchange for instance to inject calendar entries, ERP systems etc) This application should be able to export its data (from its DB) to files. I don't know if it's possible, but it would be nice if the app could add these files to MOSS and attach authorization info to them so only users with sufficient rights would be able to view/open these files. Why Asp.net MVC then? I'm very well versed in Asp.net MVC (also with the latest version) and I haven't done anything on Sharepoint since version 2003 (which doesn't do me no good or prepare me for the latest version in any way shape or form). This project will most probably be a death march project so I would rather write my application as a UI rich Asp.net MVC application and somehow integrate it into MOSS. But not only via a link, because I would like to at least share credentials, so users wouldn't need to re-login when accessing my app. Using Asp.net MVC I would at least have the possibility to finish on time or be less death marching. Is this at all possible? Questions Is it possible to integrate Asp.net MVC into MOSS as described above? If integration is not possible, would it be possible to create a completely MOSS based application that would work as described? Which parts of MOSS 2010 should I use to accomplish what I need?

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  • Scale background image to wrap content of layout

    - by bjg222
    I have a layout that contains some text fields and has a background image that's displayed at the top of my activity. I'd like the background image to scale to wrap the content (don't care about aspect ratio). However, the image is larger than content, so the layout instead wraps the background image. Here's my original code: <RelativeLayout android:layout_width="fill_parent" android:id="@+id/HeaderList" android:layout_gravity="top" android:layout_height="wrap_content" android:background="@drawable/header"> <TextView android:layout_height="wrap_content" android:layout_width="wrap_content" android:id="@+id/NameText" android:text="Jhn Doe" android:textColor="#FFFFFF" android:textSize="30sp" android:layout_alignParentLeft="true" android:layout_alignParentTop="true" android:paddingLeft="4dp" android:paddingTop="4dp" /> <TextView android:layout_height="wrap_content" android:layout_width="wrap_content" android:textColor="#FFFFFF" android:layout_alignParentLeft="true" android:id="@+id/HoursText" android:text="170 hours" android:textSize="23sp" android:layout_below="@+id/NameText" android:paddingLeft="4dp" /> </RelativeLayout> After searching through some other questions, I found these two: How to wrap content views rather than background drawable? Scale a Drawable or background image? Based on this, I created a FrameLayout w/ an ImageView showing the background. Unfortunately, I still can't get it to work. I want the height of the background image to shrink/expand w/ the size of the text views, but with the FrameLayout, the ImageView fits to the size of it's parent, and I can't find a way to make the parent fit to the size the text view layout. Here's my updated code: <FrameLayout android:layout_width="fill_parent" android:layout_height="wrap_content" > <ImageView android:src="@drawable/header" android:layout_width="fill_parent" android:scaleType="fitXY" android:layout_height="fill_parent" /> <RelativeLayout android:layout_width="fill_parent" android:id="@+id/HeaderList" android:layout_gravity="top" android:layout_height="wrap_content" > <TextView android:layout_height="wrap_content" android:layout_width="wrap_content" android:id="@+id/NameText" android:text="John Doe" android:textColor="#FFFFFF" android:textSize="30sp" android:layout_alignParentLeft="true" android:layout_alignParentTop="true" android:paddingLeft="4dp" android:paddingTop="4dp" /> <TextView android:layout_height="wrap_content" android:layout_width="wrap_content" android:textColor="#FFFFFF" android:layout_alignParentLeft="true" android:id="@+id/HoursText" android:text="170 hours" android:textSize="23sp" android:layout_below="@+id/NameText" android:paddingLeft="4dp" /> </RelativeLayout> </FrameLayout> Does anybody have any suggestions for how best to make an image scale to the size of the contents of some layout? I'm not concerned with the aspect ratio of the image, as it won't matter, I just want it to fill the background. Thanks!

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  • How to Practice Unix Programming in C?

    - by danben
    After five years of professional Java (and to a lesser extent, Python) programming and slowly feeling my CS education slip away, I decided I wanted to broaden my horizons / general usefulness to the world and do something that feels more (to me) like I really have an influence over the machine. I chose to learn C and Unix programming since I feel like that is where many of the most interesting problems are. My end goal is to be able to do this professionally, if for no other reason than the fact that I have to spend 40-50 hours per week on work that pays the bills, so it may as well also be the type of coding I want to get better at. Of course, you don't get hired to do things you haven't dont before, so for now I am ramping up on my own. To this end, I started with K&R, which was a great resource in part due to the exercises spread throughout each chapter. After that I moved on to Computer Systems: A Programmer's Perspective, followed by ten chapters of Advanced Programming in the Unix Environment. When I am done with this book, I will read Unix Network Programming. What I'm missing in the Stevens books is the lack of programming problems; they mainly document functionality and provide examples, with a few end-of-chapter questions following. I feel that I would benefit much more from being challenged to use the knowledge in each chapter ala K&R. I could write some test program for each function, but this is a less desirable method as (1) I would probably be less motivated than if I were rising to some external challenge, and (2) I will naturally only think to use the function in the ways that have already occurred to me. So, I'd like to get some recommendations on how to practice. Obviously, my first choice would be to find some resource that has Unix programming challenges. I have also considered finding and attempting to contribute to some open source C project, but this is a bit daunting as there would be some overhead in learning to use the software, then learning the codebase. The only open-source C project I can think of that I use regularly is Python, and I'm not sure how easy that would be to get started on. That said, I'm open to all kinds of suggestions as there are likely things I haven't even thought of.

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  • Problems related to showing MessageBox from non-GUI threads

    - by Hans Løken
    I'm working on a heavily data-bound Win.Forms application where I've found some strange behavior. The app has separate I/O threads receiving updates through asynchronous web-requests which it then sends to the main/GUI thread for processing and updating of application-wide data-stores (which in turn may be data-bound to various GUI-elements, etc.). The server at the other end of the web-requests requires periodic requests or the session times out. I've gone through several attempted solutions of dealing with thread-issues etc. and I've observed the following behavior: If I use Control.Invoke for sending updates from I/O-thread(s) to main-thread and this update causes a MessageBox to be shown the main form's message pump stops until the user clicks the ok-button. This also blocks the I/O-thread from continuing eventually leading to timeouts on the server. If I use Control.BeginInvoke for sending updates from I/O-thread(s) to main-thread the main form's message pump does not stop, but if the processing of an update leads to a messagebox being shown, the processing of the rest of that update is halted until the user clicks ok. Since the I/O-threads keep running and the message pump keeps processing messages several BeginInvoke's for updates may be called before the one with the message box is finished. This leads to out-of-sequence updates which is unacceptable. I/O-threads add updates to a blocking queue (very similar to http://stackoverflow.com/questions/530211/creating-a-blocking-queuet-in-net/530228#530228). GUI-thread uses a Forms.Timer that periodically applies all updates in the blocking queue. This solution solves both the problem of blocking I/O threads and sequentiality of updates i.e. next update will be never be started until previous is finished. However, there is a small performance cost as well as introducing a latency in showing updates that is unacceptable in the long run. I would like update-processing in the main-thread to be event-driven rather than polling. So to my question. How should I do this to: avoid blocking the I/O-threads guarantee that updates are finished in-sequence keep the main message pump running while showing a message box as a result of an update.

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  • Populating ComboBoxDataColumn items and values

    - by MarceloRamires
    I have a "populate combobox", and I'm so happy with it that I've even started using more comboboxes. It takes the combobox object by reference with the ID of the "value set" (or whatever you want to call it) from a table and adds the items and their respective values (which differ) and does the job. I've recently had the brilliant idea of using comboboxes in a gridview, and I was happy to notice that it worked JUST LIKE a single combobox, but populating all the comboboxes in the given column at the same time. ObjComboBox.Items.Add("yadayada"); //works just like ObjComboBoxColumn.Items.Add("blablabla"); But When I started planning how to populate these comboboxes I've noticed: There's no "Values" property in ComboBoxDataColumn. ObjComboBox.Values = whateverArray; //works, but the following doesn't ObjComboBoxColumn.Values = whateverArray; Questions: 0 - How do I populate it's values ? (I suspect it's just as simple, but uses another name) 1 - If it works just like a combobox, what's the explanation for not having this attribute ? -----[EDIT]------ So I've checked out Charles' quote, and I've figured I had to change my way of populating these bad boys. Instead of looping through the strings and inserting them one by one in the combobox, I should grab the fields I want to populate in a table, and set one column of the table as the "value", and other one as the "display". So I've done this: ObjComboBoxColumn.DataSource = DTConfig; //Double checked, guaranteed to be populated ObjComboBoxColumn.ValueMember = "Code"; ObjComboBoxColumn.DisplayMember = "Description"; But nothing happens, if I use the same object as so: ObjComboBoxColumn.Items.Add("StackOverflow"); It is added. There is no DataBind() function. It finds the two columns, and that's guaranteed ("Code" and "Description") and if I change their names to nonexistant ones it gives me an exception, so that's a good sign. -----[EDIT]------ I have a table in SQL Server that is something like code  |  text —————    1    | foo    2    | bar It's simple, and with other comboboxes (outside of gridviews) i've successfully populated looping through the rows and adding the texts: ObjComboBox.Items.Add(MyDataTable.Rows[I]["MyColumnName"].ToString()); And getting every value, adding it into an array, and setting it like: ObjComboBox.Values = MyArray; I'd like to populate my comboboxColumns just as simply as I do with comboboxes.

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