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  • How to Determine the Size of MSADO Command Parameters

    - by Adam
    I am new to MS ADO and trying to understand how to set the size on command parameters as created by the command.CreateParameter (Name, Type, Direction, Size, Value) The documentation says the following: Size Optional. A Long value that specifies the maximum length for the parameter value in characters or bytes. ... If you specify a variable-length data type in the Type argument, you must either pass a Size argument or set the Size property of the Parameter object before appending it to the Parameters collection; otherwise, an error occurs. 1.) What should one pass for fixed-size parameters? Is it a "don't care"? I was a bit confused by the example found here, in which they set size to 3 for an adInteger parameter with Value set to a variant of type VT_I2 pPrmByRoyalty->Type = adInteger; pPrmByRoyalty->Size = 3; pPrmByRoyalty->Direction = adParamInput; pPrmByRoyalty->Value = vtroyal; VT_I2 implies two bytes. A tagVARIANT struct is 16 bytes. How did they land on three? I see that the enum value for adInteger happens to be three, but I suspect that is just a coincidence. So it's a bit confusing what to pass for fixed-size parameters. The team I'm working with has always passed sizeof(int) for adInteger, and it seems to work. Is that correct? Now, for "variable-length" parameters: we are instructed by the documentation to pass "the maximum length .. in characters or bytes". 2.) For adVarChar, is it sufficient to pass the max width as defined in the database? 3.) What about the Wide types (e.g. adVarWChar)? Is it characters or bytes? 4.) How about adVariant, which could contain fixed- or variable-length data? 5.) Do arrays ever come into play here? (we don't pass them as parameters, just curious) Any references or personal insights are welcome.

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  • What am I missing about WCF?

    - by Bigtoe
    I've been developing in MS technologies for longer than I care to remember at this stage. When .NET arrived on the scene I thought they hit the nail on the head and with each iteration and version I thought their technologies were getting stronger and stronger and looked forward to each release. However, having had to work with WCF for the last year I must say I found the technology very difficult to work with and understand. Initially it's quite appealing but when you start getting into the guts of it, configuration is a nightmare, having to override behaviours for message sizes, number of objects contained in a messages, the complexity of the security model, disposing of proxies when faulted and finally moving back to defining interfaces in code rather than in XML. It just does not work out of the box and I think it should. We found all of the above issues while either testing ourselves or else when our products were out on site. I do understand the rationale behind it all, but surely they could have come up with simpler implementation mechanism. I suppose what I'm asking is, Am I looking at WCF the wrong way? What strengths does it have over the alternatives? Under what circumstances should I choose to use WCF? OK Folks, Sorry about the delay in responding, work does have a nasty habbit of get in the way somethimes :) Some clarifications My main paint point with WCF I suppose falls down into the following areas While it does work out of the box, your left with some major surprises under the hood. As pointed out above basic things are restricted until they are overridden Size of string than can be passed can't be over 8K Number of objects that can be passed in a single message is restricted Proxies not automatically recovering from failures The amount of configuration while it's there is a good thing, but understanding it all and what to use what and under which circumstances can be difficult to understand. Especially when deploying software on site with different security requirements etc. When talking about configuration, we've had to hide lots of ours in a back-end database because security and network people on-site were trying to change things in configuration files without understanding it. Keeping the configuration of the interfaces in code rather than moving to explicitly defined interfaces in XML, which can be published and consumed by almost anything. I know we can export the XML from the assembley, but it's full of rubbish and certain code generators choke on it. I know the world moves on, I've moved on a number of times over the last (ahem 22 years I've been developing) and am actively using WCF, so don't get me wrong, I do understand what it's for and where it's heading. I just think there should be simplier configuration/deployment options available, easier set-up and better management for configuration (SQL config provider maybe, rahter than just the web.config/app.config files). OK, back to the daily grid. Thanks for all your replies so far. Kind Regards Noel

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  • What do you expect from a package manager for Emacs

    - by tarsius
    Although several hundred Emacs Lisp libraries exist GNU Emacs does not have an (internal) package manager. I guess that most users would agree that it is currently rather inconvenient to find, install and especially keep up-to-date Emacs Lisp libraries. These pages make life a bit easier Emacs Lisp List - Problem: I see dead people (links). Emacswiki - Problem: May contain traces of nuts (malicious code). These are some package managers XEmacs package manager package.el - ELPA pases install.el install-elisp.el plugin.el use-package.el jem-pkg.el epkg/elm - the one I am working. And this are some packages that provide functionality that might be useful in a package manager ell.el - Browse the Emacs Lisp List genauto.el - helps generate autoloads for your elisp packages date-calc.el - date calculation and parsing routines strptime.el - partial implementation of POSIX date and time parsing wikirel.el - Visit relevant pages on the Emacs Wiki loadhist.el, lib-requires.el, elisp-depend.el - Commands to list Emacs Lisp library dependencies. project-root.el - Define a project root and take actions based upon it So I would like to know from you what you consider important/unimportant/supplementary... in a package manager for Emacs. Some ideas Many packages (incorporate the Emacs Lisp List and other lists of libraries). Only packages that have been tested. Support for more than one package archive (so people can choose between many/tested packages). Dependency calculated based on required features only. Dependencies take particular versions into account. Only use versions that have been released upstream. Use versions from version control systems if available. Packages are categorized. Packages can be uninstalled and updated not only installed. Support creating fork of upstream version of packages. Support publishing these forks. Support choosing a fork. After installation packages are activated. Generate autoloads. Integration with Emacswiki (see wikirel.el). Users can tag, comment ... packages and share that information. Only FSF-assigned/GPL/FOSS software or don't care about license. Package manager should be integrated in Emacs. Support contacting author. Lots of metadata. Suggest alternatives before installing a particular package. Some discussions about the subject at hand emacs-devel 20080801 comp.emacs 20021121 RationalElispPackaging I am hoping for these kinds of answers Pointers to more implementations, discussions etc. Lengthy descriptions of a set of features that make up your ideal package manager. Descriptions of one particular disired/undisired feature. This has the advantage that the regular voting mechanism allows us to see what features are most welcomed. Feel free to elaborate on my ideas from above. Surprise me.

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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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  • Use a subdirectory as root with htaccess in Apache 1.3

    - by Andrew
    I'm trying to deploy a site generated with Jekyll and would like to keep the site in its own subfolder on my server to keep everything more organized. Essentially, I'd like to use the contents of /jekyll as the root unless a file similarly named exists in the actual web root. So something like /jekyll/sample-page/ would show as http://www.example.com/sample-page/, while something like /other-folder/ would display as http://www.example.com/other-folder. My test server runs Apache 2.2 and the following .htaccess (adapted from http://gist.github.com/97822) works flawlessly: RewriteEngine On # Map http://www.example.com to /jekyll. RewriteRule ^$ /jekyll/ [L] # Map http://www.example.com/x to /jekyll/x unless there is a x in the web root. RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteCond %{REQUEST_URI} !^/jekyll/ RewriteRule ^(.*)$ /jekyll/$1 # Add trailing slash to directories without them so DirectoryIndex works. # This does not expose the internal URL. RewriteCond %{REQUEST_FILENAME} -d RewriteCond %{REQUEST_FILENAME} !/$ RewriteRule ^(.*)$ $1/ # Disable auto-adding slashes to directories without them, since this happens # after mod_rewrite and exposes the rewritten internal URL, e.g. turning # http://www.example.com/about into http://www.example.com/jekyll/about. DirectorySlash off However, my production server runs Apache 1.3, which doesn't allow DirectorySlash. If I disable it, the server gives a 500 error because of internal redirect overload. If I comment out the last section of ReWriteConds and rules: RewriteCond %{REQUEST_FILENAME} -d RewriteCond %{REQUEST_FILENAME} !/$ RewriteRule ^(.*)$ $1/ …everything mostly works: http://www.example.com/sample-page/ displays the correct content. However, if I omit the trailing slash, the URL in the address bar exposes the real internal URL structure: http://www.example.com/jekyll/sample-page/ What is the best way to account for directory slashes in Apache 1.3, where useful tools like DirectorySlash don't exist? How can I use the /jekyll/ directory as the site root without revealing the actual URL structure? Edit: After a ton of research into Apache 1.3, I've found that this problem is essentially a combination of two different issues listed at the Apache 1.3 URL Rewriting Guide. I have a (partially) moved DocumentRoot, which in theory would be taken care of with something like this: RewriteRule ^/$ /e/www/ [R] I also have the infamous "Trailing Slash Problem," which is solved by setting the RewriteBase (as was suggested in one of the responses below): RewriteBase /~quux/ RewriteRule ^foo$ foo/ [R] The problem is combining the two. Moving the document root doesn't (can't?) use RewriteBase—fixing trailing slashes requires(?) it… Hmm…

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  • Java: error handling with try-catch, empty-try-catch, dummy-return

    - by HH
    A searh uses recursively defined function that easily throws exceptions. I have tried 3 ways to handle exeptions: to ignore with an empty-try-catch() add-dummy-return stop err-propagation due to exeption throw a specific except. (this part I don't really understand. If I throw except, can I force it to continue elsewhere, not continuing the old except-thrown-path?) Some exceptions I do not realy care like during execution removed files -exception (NullPointer) but some I really do like unknown things. Possible exceptions: // 1. if a temp-file or some other file removed during execution -> except. // 2. if no permiss. -> except. // 3. ? --> except. The code is Very import for the whole program. I earlier added clittered-checks, try-catches, avoided-empty-try-catches but it really blurred the logic. Some stoned result here would make the code later much easier to maintain. It was annoying to track random exeptions due to some random temp-file removal! How would you handle exceptions for the critical part? Code public class Find { private Stack<File> fs=new Stack<File>(); private Stack<File> ds=new Stack<File>(); public Stack<File> getD(){ return ds;} public Stack<File> getF(){ return fs;} public Find(String path) { // setting this type of special checks due to errs // propagation makes the code clittered if(path==null) { System.out.println("NULL in Find(path)"); System.exit(9); } this.walk(path); } private void walk( String path ) { File root = new File( path ); File[] list = root.listFiles(); //TODO: dangerous with empty try-catch?! try{ for ( File f : list ) { if ( f.isDirectory() ) { walk( f.getAbsolutePath() ); ds.push(f); } else { fs.push(f); } } }catch(Exception e){e.printStackTrace();} } } Code refactored from here.

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  • heroku time zone problem, logging local server time

    - by Ole Morten Amundsen
    UPDATE: Ok, I didn't formulate a good Q to be answered. I still struggle with heroku being on -07:00 UTC and I at +02:200 UTC. Q: How do I get the log written in the correct Time.zone ? The 9 hours difference, heroku (us west) - norway, is distracting to work with. I get this in my production.log (using heroku logs): Processing ProductionController#create to xml (for 81.26.51.35 at 2010-04-28 23:00:12) [POST] How do I get it to write 2010-04-29 08:00:12 +02:00 GMT ? Note that I'm running at heroku and cannot set the server time myself, as one could do at your amazon EC2 servers. Below is my previous question, I'll leave it be as it holds some interesting information about time and zones. Why does Time.now yield the server local time when I have set the another time zone in my environment.rb config.time_zone = 'Copenhagen' I've put this in a view <p> Time.zone <%= Time.zone %> </p> <p> Time.now <%= Time.now %> </p> <p> Time.now.utc <%= Time.now.utc %> </p> <p> Time.zone.now <%= Time.zone.now %> </p> <p> Time.zone.today <%= Time.zone.today %> </p> rendering this result on my app at heroku Time.zone (GMT+01:00) Copenhagen Time.now Mon Apr 26 08:28:21 -0700 2010 Time.now.utc Mon Apr 26 15:28:21 UTC 2010 Time.zone.now 2010-04-26 17:28:21 +0200 Time.zone.today 2010-04-26 Time.zone.now yields the correct result. Do I have to switch from Time.now to Time.zone.now, everywhere? Seems cumbersome. I truly don't care what the local time of the server is, it's giving me loads of trouble due to extensive use of Time.now. Am I misunderstanding anything fundamental here?

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  • LSP packet modify

    - by kellogs
    Hello, anybody care to share some insights on how to use LSP for packet modifying ? I am using the non IFS subtype and I can see how (pseudo?) packets first enter WSPRecv. But how do I modify them ? My inquiry is about one single HTTP response that causes WSPRecv to be called 3 times :((. I need to modify several parts of this response, but since it comes in 3 slices, it is pretty hard to modify it accordingly. And, maybe on other machines or under different conditions (such as high traffic) there would only be one sole WSPRecv call, or maybe 10 calls. What is the best way to work arround this (please no NDIS :D), and how to properly change the buffer (lpBuffers-buf) by increasing it ? int WSPAPI WSPRecv( SOCKET s, LPWSABUF lpBuffers, DWORD dwBufferCount, LPDWORD lpNumberOfBytesRecvd, LPDWORD lpFlags, LPWSAOVERLAPPED lpOverlapped, LPWSAOVERLAPPED_COMPLETION_ROUTINE lpCompletionRoutine, LPWSATHREADID lpThreadId, LPINT lpErrno ) { LPWSAOVERLAPPEDPLUS ProviderOverlapped = NULL; SOCK_INFO *SocketContext = NULL; int ret = SOCKET_ERROR; *lpErrno = NO_ERROR; // // Find our provider socket corresponding to this one // SocketContext = FindAndRefSocketContext(s, lpErrno); if ( NULL == SocketContext ) { dbgprint( "WSPRecv: FindAndRefSocketContext failed!" ); goto cleanup; } // // Check for overlapped I/O // if ( NULL != lpOverlapped ) { /*bla bla .. not interesting in my case*/ } else { ASSERT( SocketContext->Provider->NextProcTable.lpWSPRecv ); SetBlockingProvider(SocketContext->Provider); ret = SocketContext->Provider->NextProcTable.lpWSPRecv( SocketContext->ProviderSocket, lpBuffers, dwBufferCount, lpNumberOfBytesRecvd, lpFlags, lpOverlapped, lpCompletionRoutine, lpThreadId, lpErrno); SetBlockingProvider(NULL); //is this the place to modify packet length and contents ? if (strstr(lpBuffers->buf, "var mapObj = null;")) { int nLen = strlen(lpBuffers->buf) + 200; /*CHAR *szNewBuf = new CHAR[]; CHAR *pIndex; pIndex = strstr(lpBuffers->buf, "var mapObj = null;"); nLen = strlen(strncpy(szNewBuf, lpBuffers->buf, (pIndex - lpBuffers->buf) * sizeof (CHAR))); nLen = strlen(strncpy(szNewBuf + nLen * sizeof(CHAR), "var com = null;\r\n", 17 * sizeof(CHAR))); pIndex += 18 * sizeof(CHAR); nLen = strlen(strncpy(szNewBuf + nLen * sizeof(CHAR), pIndex, 1330 * sizeof (CHAR))); nLen = strlen(strncpy(szNewBuf + nLen * sizeof(CHAR), "if (com == null)\r\n" \ "com = new ActiveXObject(\"InterCommJS.Gateway\");\r\n" \ "com.lat = latitude;\r\n" \ "com.lon = longitude;\r\n}", 111 * sizeof (CHAR))); pIndex = strstr(szNewBuf, "Content-Length:"); pIndex += 16 * sizeof(CHAR); strncpy(pIndex, "1465", 4 * sizeof(CHAR)); lpBuffers->buf = szNewBuf; lpBuffers->len += 128;*/ } if ( SOCKET_ERROR != ret ) { SocketContext->BytesRecv += *lpNumberOfBytesRecvd; } } cleanup: if ( NULL != SocketContext ) DerefSocketContext( SocketContext, lpErrno ); return ret; } Thank you

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  • Cases of companies taking IP rights of your own personal projects developed outside company time

    - by GSS
    Hi, I have heard of cases where a developer working for a company is also making his own personal projects in his own time, using his own equipment yet the company he works for tries to claim ownership for the project. I really find this annoying, and bang out of order. It should also be illegal. I am in this position (work for a company and working on my own systems - from small class libraries used to practise what I learn in my exam revision to a large commercial-scale system). While I don't know if the company will try to take ownership, all I know is they say they do not want a conflict of interest. Fair enough, my system is developed in my own time using my own equipment. They also say that work time should be for work only, which it is. Funny thing that as work is so boring, easy and slow that I have plenty of free time, which I wish I could spend on something productive - said system. The problem is, my company does not take hiring technical talent seriously. This is my first job, I am a junior coder (but my status/position doesn't really reflect what I can do), but I am the only developer. Likewise with the guy who controls Windows Server. As the contract does not say anything about taking ownership, I would assume they would. They would try to milk my success (I've made a good impression so I am sure they would). How can this be allowed? Are there any examples of this happening to any fellow Stacker here? It really makes my blood boil. What I find funny is that my company hardly has the expertise and resources to even be able to successfully run a project of my size. What I do at work is an ASP.NET application consisting of five pages, and even then there are flaws in the project. If I told them that they would also have to take responsibility for flaws in the project, then they would think twice! It's exactly because of this I save the best code for myself and at work I write rubbish code full of code smells. The company don't really care about error handling, as long as the business functionality works (ie a scheduled email sends, but there is no error handling). They'd think twice when they see the embarassment and business cost of a YSOD...

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  • Advice on logic circuits and serial communications

    - by Spencer Ruport
    As far as I understand the serial port so far, transferring data is done over pin 3. As shown here: There are two things that make me uncomfortable about this. The first is that it seems to imply that the two connected devices agree on a signal speed and the second is that even if they are configured to run at the same speed you run into possible synchronization issues... right? Such things can be handled I suppose but it seems like there must be a simpler method. What seems like a better approach to me would be to have one of the serial port pins send a pulse that indicates that the next bit is ready to be stored. So if we're hooking these pins up to a shift register we basically have: (some pulse pin)-clk, tx-d Is this a common practice? Is there some reason not to do this? EDIT Mike shouldn't have deleted his answer. This I2C (2 pin serial) approach seems fairly close to what I did. The serial port doesn't have a clock you're right nobugz but that's basically what I've done. See here: private void SendBytes(byte[] data) { int baudRate = 0; int byteToSend = 0; int bitToSend = 0; byte bitmask = 0; byte[] trigger = new byte[1]; trigger[0] = 0; SerialPort p; try { p = new SerialPort(cmbPorts.Text); } catch { return; } if (!int.TryParse(txtBaudRate.Text, out baudRate)) return; if (baudRate < 100) return; p.BaudRate = baudRate; for (int index = 0; index < data.Length * 8; index++) { byteToSend = (int)(index / 8); bitToSend = index - (byteToSend * 8); bitmask = (byte)System.Math.Pow(2, bitToSend); p.Open(); p.Parity = Parity.Space; p.RtsEnable = (byte)(data[byteToSend] & bitmask) > 0; s = p.BaseStream; s.WriteByte(trigger[0]); p.Close(); } } Before anyone tells me how ugly this is or how I'm destroying my transfer speeds my quick answer is I don't care about that. My point is this seems much much simpler than the method you described in your answer nobugz. And it wouldn't be as ugly if the .Net SerialPort class gave me more control over the pin signals. Are there other serial port APIs that do?

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  • Java Hardware Acceleration

    - by Freezerburn
    I have been spending some time looking into the hardware acceleration features of Java, and I am still a bit confused as none of the sites that I found online directly and clearly answered some of the questions I have. So here are the questions I have for hardware acceleration in Java: 1) In Eclipse version 3.6.0, with the most recent Java update for Mac OS X (1.6u10 I think), is hardware acceleration enabled by default? I read somewhere that someCanvas.getGraphicsConfiguration().getBufferCapabilities().isPageFlipping() is supposed to give an indication of whether or not hardware acceleration is enabled, and my program reports back true when that is run on my main Canvas instance for drawing to. If my hardware acceleration is not enabled now, or by default, what would I have to do to enable it? 2) I have seen a couple articles here and there about the difference between a BufferedImage and VolatileImage, mainly saying that VolatileImage is the hardware accelerated image and is stored in VRAM for fast copy-from operations. However, I have also found some instances where BufferedImage is said to be hardware accelerated as well. Is BufferedImage hardware accelerated as well in my environment? What would be the advantage of using a VolatileImage if both types are hardware accelerated? My main assumption for the advantage of having a VolatileImage in the case of both having acceleration is that VolatileImage is able to detect when its VRAM has been dumped. But if BufferedImage also support acceleration now, would it not have the same kind of detection built into it as well, just hidden from the user, in case that the memory is dumped? 3) Is there any advantage to using someGraphicsConfiguration.getCompatibleImage/getCompatibleVolatileImage() as opposed to ImageIO.read() In a tutorial I have been reading for some general concepts about setting up the rendering window properly (tutorial) it uses the getCompatibleImage method, which I believe returns a BufferedImage, to get their "hardware accelerated" images for fast drawing, which ties into question 2 about if it is hardware accelerated. 4) This is less hardware acceleration, but it is something I have been curious about: do I need to order which graphics get drawn? I know that when using OpenGL via C/C++ it is best to make sure that the same graphic is drawn in all the locations it needs to be drawn at once to reduce the number of times the current texture needs to be switch. From what I have read, it seems as if Java will take care of this for me and make sure things are drawn in the most optimal fashion, but again, nothing has ever said anything like this clearly. 5) What AWT/Swing classes support hardware acceleration, and which ones should be used? I am currently using a class that extends JFrame to create a window, and adding a Canvas to it from which I create a BufferStrategy. Is this good practice, or is there some other type of way I should be implementing this? Thank you very much for your time, and I hope I provided clear questions and enough information for you to answer my several questions.

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  • What are five things you hate about your favorite language?

    - by brian d foy
    There's been a cluster of Perl-hate on Stackoverflow lately, so I thought I'd bring my "Five things you hate about your favorite language" question to StackOverflow. Take your favorite language and tell me five things you hate about it. Those might be things that just annoy you, admitted design flaws, recognized performance problems, or any other category. You just have to hate it, and it has to be your favorite language. Don't compare it to another language, and don't talk about languages that you already hate. Don't talk about the things you like in your favorite language. I just want to hear the things that you hate but tolerate so you can use all of the other stuff, and I want to hear it about the language you wished other people would use. I ask this whenever someone tries to push their favorite language on me, and sometimes as an interview question. If someone can't find five things to hate about his favorite tool, he don't know it well enough to either advocate it or pull in the big dollars using it. He hasn't used it in enough different situations to fully explore it. He's advocating it as a culture or religion, which means that if I don't choose his favorite technology, I'm wrong. I don't care that much which language you use. Don't want to use a particular language? Then don't. You go through due diligence to make an informed choice and still don't use it? Fine. Sometimes the right answer is "You have a strong programming team with good practices and a lot of experience in Bar. Changing to Foo would be stupid." This is a good question for code reviews too. People who really know a codebase will have all sorts of suggestions for it, and those who don't know it so well have non-specific complaints. I ask things like "If you could start over on this project, what would you do differently?" In this fantasy land, users and programmers get to complain about anything and everything they don't like. "I want a better interface", "I want to separate the model from the view", "I'd use this module instead of this other one", "I'd rename this set of methods", or whatever they really don't like about the current situation. That's how I get a handle on how much a particular developer knows about the codebase. It's also a clue about how much of the programmer's ego is tied up in what he's telling me. Hate isn't the only dimension of figuring out how much people know, but I've found it to be a pretty good one. The things that they hate also give me a clue how well they are thinking about the subject.

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  • GCC, -O2, and bitfields - is this a bug or a feature?

    - by Rooke
    Today I discovered alarming behavior when experimenting with bit fields. For the sake of discussion and simplicity, here's an example program: #include <stdio.h> struct Node { int a:16 __attribute__ ((packed)); int b:16 __attribute__ ((packed)); unsigned int c:27 __attribute__ ((packed)); unsigned int d:3 __attribute__ ((packed)); unsigned int e:2 __attribute__ ((packed)); }; int main (int argc, char *argv[]) { Node n; n.a = 12345; n.b = -23456; n.c = 0x7ffffff; n.d = 0x7; n.e = 0x3; printf("3-bit field cast to int: %d\n",(int)n.d); n.d++; printf("3-bit field cast to int: %d\n",(int)n.d); } The program is purposely causing the 3-bit bit-field to overflow. Here's the (correct) output when compiled using "g++ -O0": 3-bit field cast to int: 7 3-bit field cast to int: 0 Here's the output when compiled using "g++ -O2" (and -O3): 3-bit field cast to int: 7 3-bit field cast to int: 8 Checking the assembly of the latter example, I found this: movl $7, %esi movl $.LC1, %edi xorl %eax, %eax call printf movl $8, %esi movl $.LC1, %edi xorl %eax, %eax call printf xorl %eax, %eax addq $8, %rsp The optimizations have just inserted "8", assuming 7+1=8 when in fact the number overflows and is zero. Fortunately the code I care about doesn't overflow as far as I know, but this situation scares me - is this a known bug, a feature, or is this expected behavior? When can I expect gcc to be right about this? Edit (re: signed/unsigned) : It's being treated as unsigned because it's declared as unsigned. Declaring it as int you get the output (with O0): 3-bit field cast to int: -1 3-bit field cast to int: 0 An even funnier thing happens with -O2 in this case: 3-bit field cast to int: 7 3-bit field cast to int: 8 I admit that attribute is a fishy thing to use; in this case it's a difference in optimization settings I'm concerned about.

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  • mysql index optimization for a table with multiple indexes that index some of the same columns

    - by Sean
    I have a table that stores some basic data about visitor sessions on third party web sites. This is its structure: id, site_id, unixtime, unixtime_last, ip_address, uid There are four indexes: id, site_id/unixtime, site_id/ip_address, and site_id/uid There are many different types of ways that we query this table, and all of them are specific to the site_id. The index with unixtime is used to display the list of visitors for a given date or time range. The other two are used to find all visits from an IP address or a "uid" (a unique cookie value created for each visitor), as well as determining if this is a new visitor or a returning visitor. Obviously storing site_id inside 3 indexes is inefficient for both write speed and storage, but I see no way around it, since I need to be able to quickly query this data for a given specific site_id. Any ideas on making this more efficient? I don't really understand B-trees besides some very basic stuff, but it's more efficient to have the left-most column of an index be the one with the least variance - correct? Because I considered having the site_id being the second column of the index for both ip_address and uid but I think that would make the index less efficient since the IP and UID are going to vary more than the site ID will, because we only have about 8000 unique sites per database server, but millions of unique visitors across all ~8000 sites on a daily basis. I've also considered removing site_id from the IP and UID indexes completely, since the chances of the same visitor going to multiple sites that share the same database server are quite small, but in cases where this does happen, I fear it could be quite slow to determine if this is a new visitor to this site_id or not. The query would be something like: select id from sessions where uid = 'value' and site_id = 123 limit 1 ... so if this visitor had visited this site before, it would only need to find one row with this site_id before it stopped. This wouldn't be super fast necessarily, but acceptably fast. But say we have a site that gets 500,000 visitors a day, and a particular visitor loves this site and goes there 10 times a day. Now they happen to hit another site on the same database server for the first time. The above query could take quite a long time to search through all of the potentially thousands of rows for this UID, scattered all over the disk, since it wouldn't be finding one for this site ID. Any insight on making this as efficient as possible would be appreciated :) Update - this is a MyISAM table with MySQL 5.0. My concerns are both with performance as well as storage space. This table is both read and write heavy. If I had to choose between performance and storage, my biggest concern is performance - but both are important. We use memcached heavily in all areas of our service, but that's not an excuse to not care about the database design. I want the database to be as efficient as possible.

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  • Rails "NoMethodError" with sub-resources

    - by Tchock
    Hi. I'm a newbie Rails developer who is getting the following error when trying to access the 'new' action on my CityController: undefined method `cities_path' for #<#<Class:0x104608c18>:0x104606f08> Extracted source (around line #2): 1: <h1>New City</h1> 2: <%= form_for(@city) do |f| %> 3: <%= f.error_messages %> 4: 5: <div class="field"> As some background, I have a State model with many Cities. I'm getting this error after clicking on the following link coming from a State show page: <p>Add a city: <%= link_to "Add city", new_state_city_path(@state) %></p> When I run 'rake:routes' it says this is a legit route... For more background, here is the CityController 'new' action: def new @city = City.new respond_to do |format| format.html # new.html.erb format.xml { render :xml => @city } end end Here is the (complete) form in the view: <%= form_for(@city) do |f| %> <%= f.error_messages %> <div class="field"> <%= f.label :name %><br /> <%= f.text_field :name %> </div> <div class="actions"> <%= f.submit %> </div> <% end %> This initially made me think that it's a resources/routes issue since it came back with a mention of 'cities_path' (in fact, that's what another person posting to Stack Overflow had wrong (http://stackoverflow.com/questions/845315/rails-error-nomethoderror-my-first-ruby-app). However, that doesn't seem to be the case from what I can see. Here are how my resources look in my routes file: resources :states do resources :cities end I can get it working when they are not sub-resources, but I really need to keep them as sub-resources for my future plans with the app. Any help would be very much appreciated, since I've been racking my brains on this for more hours than I would care to admit... Thanks! (Not sure this matters at all, but I'm running the very latest version of Rails 3 beta2).

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  • Applying Domain Model on top of Linq2Sql entities

    - by Thomas
    I am trying to practice the model first approach and I am putting together a domain model. My requirement is pretty simple: UserSession can have multiple ShoppingCartItems. I should start off by saying that I am going to apply the domain model interfaces to Linq2Sql generated entities (using partial classes). My requirement translates into three database tables (UserSession, Product, ShoppingCartItem where ProductId and UserSessionId are foreign keys in the ShoppingCartItem table). Linq2Sql generates these entities for me. I know I shouldn't even be dealing with the database at this point but I think it is important to mention. The aggregate root is UserSession as a ShoppingCartItem can not exist without a UserSession but I am unclear on the rest. What about Product? It is defiently an entity but should it be associated to ShoppingCartItem? Here are a few suggestion (they might all be incorrect implementations): public interface IUserSession { public Guid Id { get; set; } public IList<IShoppingCartItem> ShoppingCartItems{ get; set; } } public interface IShoppingCartItem { public Guid UserSessionId { get; set; } public int ProductId { get; set; } } Another one would be: public interface IUserSession { public Guid Id { get; set; } public IList<IShoppingCartItem> ShoppingCartItems{ get; set; } } public interface IShoppingCartItem { public Guid UserSessionId { get; set; } public IProduct Product { get; set; } } A third one is: public interface IUserSession { public Guid Id { get; set; } public IList<IShoppingCartItemColletion> ShoppingCartItems{ get; set; } } public interface IShoppingCartItemColletion { public IUserSession UserSession { get; set; } public IProduct Product { get; set; } } public interface IProduct { public int ProductId { get; set; } } I have a feeling my mind is too tightly coupled with database models and tables which is making this hard to grasp. Anyone care to decouple?

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  • A simple Python deployment problem - a whole world of pain

    - by Evgeny
    We have several Python 2.6 applications running on Linux. Some of them are Pylons web applications, others are simply long-running processes that we run from the command line using nohup. We're also using virtualenv, both in development and in production. What is the best way to deploy these applications to a production server? In development we simply get the source tree into any directory, set up a virtualenv and run - easy enough. We could do the same in production and perhaps that really is the most practical solution, but it just feels a bit wrong to run svn update in production. We've also tried fab, but it just never works first time. For every application something else goes wrong. It strikes me that the whole process is just too hard, given that what we're trying to achieve is fundamentally very simple. Here's what we want from a deployment process. We should be able to run one simple command to deploy an updated version of an application. (If the initial deployment involves a bit of extra complexity that's fine.) When we run this command it should copy certain files, either out of a Subversion repository or out of a local working copy, to a specified "environment" on the server, which probably means a different virtualenv. We have both staging and production version of the applications on the same server, so they need to somehow be kept separate. If it installs into site-packages, that's fine too, as long as it works. We have some configuration files on the server that should be preserved (ie. not overwritten or deleted by the deployment process). Some of these applications import modules from other applications, so they need to be able to reference each other as packages somehow. This is the part we've had the most trouble with! I don't care whether it works via relative imports, site-packages or whatever, as long as it works reliably in both development and production. Ideally the deployment process should automatically install external packages that our applications depend on (eg. psycopg2). That's really it! How hard can it be?

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  • MVC 2 AntiForgeryToken - Why symmetric encryption + IPrinciple?

    - by Brad R
    We recently updated our solution to MVC 2, and this has updated the way that the AntiForgeryToken works. Unfortunately this does not fit with our AJAX framework any more. The problem is that MVC 2 now uses symmetric encryption to encode some properties about the user, including the user's Name property (from IPrincipal). We are able to securely register a new user using AJAX, after which subsequent AJAX calls will be invalid as the anti forgery token will change when the user has been granted a new principal. There are also other cases when this may happen, such as a user updating their name etc. My main question is why does MVC 2 even bother using symmetric encryption? Any then why does it care about the user name property on the principal? If my understanding is correct then any random shared secret will do. The basic principle is that the user will be sent a cookie with some specific data (HttpOnly!). This cookie is then required to match a form variable sent back with each request that may have side effects (POST's usually). Since this is only meant to protect from cross site attacks it is easy to craft up a response that would easily pass the test, but only if you had full access to the cookie. Since a cross site attacker is not going to have access to your user cookies you are protected. By using symmetric encryption, what is the advantage in checking the contents of the cookie? That is, if I already have sent an HttpOnly cookie the attacker cannot override it (unless a browser has a major security issue), so why do I then need to check it again? After having a think about it it appears to be one of those 'added layer of security' cases - but if your first line of defence has fallen (HttpOnly) then the attacker is going to get past the second layer anyway as they have full access to the users cookie collection, and could just impersonate them directly, instead of using an indirect XSS/CSRF attack. Of course I could be missing a major issue, but I haven't found it yet. If there are some obvious or subtle issues at play here then I would like to be aware of them.

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  • Issues with Rails 3.1 API with Query String to Create action on Mac OSX Mountain Lion

    - by hjaved
    Hi I've been stuck on this problem for a while and would appreciate your help. I'm writing an API to allow an external source like a Browser Query String or a smartphone to enter some model User info in a form and hit the User create action to write the data to the db. Please tell me what I'm doing wrong with the code below. I've also observed that if I have code like @user = User.new(params[:user]), that this approach only works when a user enters their data within the form. And that if I have code such as @user = User.new( name: params[:name], location: params[:location], password = params[:password], email: params[:email]), that this code ONLY works for a Query string entry, but NOT both Query string AND regular form submission. Why is that and how can I write the code above in the Users Controller Create action, so that it takes care of both situations? URL used: localhost:3000/users/create?name=John&&[email protected]&&password=secret&&location=SanFrancisco&date=06122012 The date is of type string but it doesn't show up in the database. Why? Everything else does. UsersController.rb def create @user = User.new(params[:user]) if @user.save session[:uid] = @user.id redirect_to thanks_path, notice: "Welcome #{@user.name}!" else redirect_to root_path end end New User Form: <%=u.text_field :name, placeholder: "Name"%><br> <%=u.text_field :email, placeholder: "Email"%><br> <%=u.password_field :password, placeholder: "Password"%><br> <%=u.text_field :location, placeholder: "City"%><br> <%=u.text_field :date, placeholder: "Date"%><br> <%if params[:partner_id]%> <%=u.hidden_field :partner_id, value: params[:partner_id]%> <%end%> <button class="btn btn-large btn-primary">Enter</button> I also tried to create a separate method called remotecreate for User creation for something other than a regular web form. I entered remotecreate in the Query string but it didn't work. def remotecreate @user = User.create(name: params[:name], email: params[:email], password: params[:password], location: params[:location], date: params[:date]) if @user.save session[:uid] = @user.id redirect_to thanks_path, notice: "Welcome #{@user.name}" else redirect_to root_path end end Thanks!

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  • How can I inject Javascript (including Prototype.js) in other sites without cluttering the global na

    - by Daniel Magliola
    I'm currently on a project that is a big site that uses the Prototype library, and there is already a humongous amount of Javascript code. We're now working on a piece of code that will get "injected" into other people's sites (picture people adding a <script> tag in their sites) which will then run our code and add a bunch of DOM elements and functionality to their site. This will have new pieces of code, and will also reuse a lot of the code that we use on our main site. The problem I have is that it's of course not cool to just add a <script> that will include Prototype in people's pages. If we do that in a page that's already using ANY framework, we're guaranteed to screw everything up. jQuery gives us the option to "rename" the $ object, so it could handle this situation decently, except obviously for the fact that we're not using jQuery, so we'd have to migrate everything. Right now i'm contemplating a number of ugly choices, and I'm not sure what's best... Rewrite everything to use jQuery, with a renamed $ object everywhere. Creating a "new" Prototype library with only the subset we'd be using in "injected" code, and renaming $ to something else. Then again I'd have to adapt the parts of my code that would be shared somehow. Not using a library at all in injected code, to keep it as clean as possible, and rewriting the shared code to use no library at all. This would obviously degenerate into us creating our own frankenstein of a library, which is probably the worst case scenario ever. I'm wondering what you guys think I could do, and also whether there's some magic option that would solve all my problems... For example, do you think I could use something like Caja / Cajita to sandbox my own code and isolate it from the rest of the site, and have Prototype inside of there? Or am I completely missing the point with that? I also read once about a technique for bookmarklets, were you add your code like this: (function() { /* your code */ })(); And then your code is all inside your anonymous function and you haven't touched the global namespace at all. Do you think I could make one file containing: (function() { /* Full Code of the Prototype file here */ /* All my code that will run in the "other" site */ InitializeStuff_CreateDOMElements_AttachEventHandlers(); })(); Would that work? Would it accomplish the objective of not cluttering the global namespace, and not killing the functionality on a site that uses jQuery, for example? Or is Prototype too complex somehow to isolate it like that? (NOTE: I think I know that that would create closures everywhere and that's slower, but I don't care too much about performance, my code is not doing anything that complex)

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  • i want to have some cross browser consistency on my fieldsets, do you know how can i do it?

    - by Omar
    i have this problem with fieldsets... have a look at http://i.imgur.com/IRrXB.png is it possible to achieve what i want with css??? believe me, i tried! as you can see on the img, i just want the look of the legend to be consistent across browsers, i want it to use the width of the fieldset no more (like chrome and ie) no less (like firefox), dont worry about the rounded corners and other issues, thats taken care of. heres the the core i'm using. CSS <style type="text/css"> fieldset {margin: 0 0 10px 0;padding: 0; border:1px solid silver; background-color: #f9f9f9; -moz-border-radius:5px; -webkit-border-radius:5px; border-radius:5px} fieldset p{clear:both;margin:.3em 0;overflow:hidden;} fieldset label{float:left;width:140px;display:block;text-align:right;padding-right:8px;margin-right: 2px;color: #4a4a4a;} fieldset input, fieldset textarea {margin:0;border:1px solid #ddd;padding:3px 5px 3px 5px;} fieldset legend { background: #C6D1E8; position:relative; left: -1px; margin: 0; width: 100%; padding: 0px 5px; font-size: 1.11em; font-weight: bold; text-align:left; border: 1px solid silver; -webkit-border-top-left-radius: 5px; -webkit-border-top-right-radius: 5px; -moz-border-radius-topleft: 5px; -moz-border-radius-topright: 3px; border-top-left-radius: 5px; border-top-right-radius: 5px; } #md {width: 400px;} </style> HTML <div id="md"> <fieldset> <legend>some title</legend> <p> <label>Login</label> <input type="text" /> </p> <p> <label>Password</label> <input type="text" /> </p> <p><label>&nbsp;</label> <input type="submit"> </p> </fieldset> </div>

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  • What's the best way to do base36 arithmetic in perl?

    - by DVK
    What's the best way to do base36 arithmetic in Perl? To be more specific, I need to be able to do the following: Operate on positive N-digit numbers in base 36 (e.g. digits are 0-9 A-Z) N is finite, say 9 Provide basic arithmetic, at the very least the following 3: Addition (A+B) Subtraction (A-B) Whole division, e.g. floor(A/B). Strictly speaking, I don't really need a base10 conversion ability - the numbers will 100% of time be in base36. So I'm quite OK if the solution does NOT implement conversion from base36 back to base10 and vice versa. I don't much care whether the solution is brute-force "convert to base 10 and back" or converting to binary, or some more elegant approach "natively" performing baseN operations (as stated above, to/from base10 conversion is not a requirement). My only 3 considerations are: It fits the minimum specifications above It's "standard". Currently we're using and old homegrown module based on base10 conversion done by hand that is buggy and sucks. I'd much rather replace that with some commonly used CPAN solution instead of re-writing my own bicycle from scratch, but I'm perfectly capable of building it if no better standard possibility exists. It must be fast-ish (though not lightning fast). Something that takes 1 second to sum up 2 9-digit base36 numbers is worse than anything I can roll on my own :) P.S. Just to provide some context in case people decide to solve my XY problem for me in addition to answering the technical question above :) We have a fairly large tree (stored in DB as a bunch of edges), and we need to superimpose order on a subset of that tree. The tree dimentions are big both depth- and breadth- wise. The tree is VERY actively updated (inserts and deletes and branch moves). This is currently done by having a second table with 3 columns: parent_vertex, child_vertex, local_order, where local_order is an 9-character string built of A-Z0-9 (e.g. base 36 number). Additional considerations: It is required that the local order is unique per child (and obviously unique per parent), Any complete re-ordering of a parent is somewhat expensive, and thus the implementation is to try and assign - for a parent with X children - the orders which are somewhat evenly distributed between 0 and 36**10-1, so that almost no tree inserts result in a full re-ordering.

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  • Permutations of Varying Size

    - by waiwai933
    I'm trying to write a function in PHP that gets all permutations of all possible sizes. I think an example would be the best way to start off: $my_array = array(1,1,2,3); Possible permutations of varying size: 1 1 // * See Note 2 3 1,1 1,2 1,3 // And so forth, for all the sets of size 2 1,1,2 1,1,3 1,2,1 // And so forth, for all the sets of size 3 1,1,2,3 1,1,3,2 // And so forth, for all the sets of size 4 Note: I don't care if there's a duplicate or not. For the purposes of this example, all future duplicates have been omitted. What I have so far in PHP: function getPermutations($my_array){ $permutation_length = 1; $keep_going = true; while($keep_going){ while($there_are_still_permutations_with_this_length){ // Generate the next permutation and return it into an array // Of course, the actual important part of the code is what I'm having trouble with. } $permutation_length++; if($permutation_length>count($my_array)){ $keep_going = false; } else{ $keep_going = true; } } return $return_array; } The closest thing I can think of is shuffling the array, picking the first n elements, seeing if it's already in the results array, and if it's not, add it in, and then stop when there are mathematically no more possible permutations for that length. But it's ugly and resource-inefficient. Any pseudocode algorithms would be greatly appreciated. Also, for super-duper (worthless) bonus points, is there a way to get just 1 permutation with the function but make it so that it doesn't have to recalculate all previous permutations to get the next? For example, I pass it a parameter 3, which means it's already done 3 permutations, and it just generates number 4 without redoing the previous 3? (Passing it the parameter is not necessary, it could keep track in a global or static). The reason I ask this is because as the array grows, so does the number of possible combinations. Suffice it to say that one small data set with only a dozen elements grows quickly into the trillions of possible combinations and I don't want to task PHP with holding trillions of permutations in its memory at once.

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  • Strange iPhone memory leak in xml parser

    - by Chris
    Update: I edited the code, but the problem persists... Hi everyone, this is my first post here - I found this place a great ressource for solving many of my questions. Normally I try my best to fix anything on my own but this time I really have no idea what goes wrong, so I hope someone can help me out. I am building an iPhone app that parses a couple of xml files using TouchXML. I have a class XMLParser, which takes care of downloading and parsing the results. I am getting memory leaks when I parse an xml file more than once with the same instance of XMLParser. Here is one of the parsing snippets (just the relevant part): for(int counter = 0; counter < [item childCount]; counter++) { CXMLNode *child = [item childAtIndex:counter]; if([[child name] isEqualToString:@"PRODUCT"]) { NSMutableDictionary *product = [[NSMutableDictionary alloc] init]; for(int j = 0; j < [child childCount]; j++) { CXMLNode *grandchild = [child childAtIndex:j]; if([[grandchild stringValue] length] > 1) { NSString *trimmedString = [[grandchild stringValue] stringByTrimmingCharactersInSet:[NSCharacterSet whitespaceAndNewlineCharacterSet]]; [product setObject:trimmedString forKey:[grandchild name]]; } } // Add product to current category array switch (categoryId) { case 0: [self.mobil addObject: product]; break; case 1: [self.allgemein addObject: product]; break; case 2: [self.besitzeIch addObject: product]; break; case 3: [self.willIch addObject: product]; break; default: break; } [product release]; } } The first time, I parse the xml no leak shows up in instruments, the next time I do so, I got a lot of leaks (NSCFString / NSCFDictionary). Instruments points me to this part inside CXMLNode.m, when I dig into a leaked object: theStringValue = [NSString stringWithUTF8String:(const char *)theXMLString]; if ( _node->type != CXMLTextKind ) xmlFree(theXMLString); } return(theStringValue); I really spent a long time and tried multiple approaches to fix this, but to no avail so far, maybe I am missing something essential? Any help is highly appreciated, thank you!

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  • LINQ aggregate left join on SQL CE

    - by P Daddy
    What I need is such a simple, easy query, it blows me away how much work I've done just trying to do it in LINQ. In T-SQL, it would be: SELECT I.InvoiceID, I.CustomerID, I.Amount AS AmountInvoiced, I.Date AS InvoiceDate, ISNULL(SUM(P.Amount), 0) AS AmountPaid, I.Amount - ISNULL(SUM(P.Amount), 0) AS AmountDue FROM Invoices I LEFT JOIN Payments P ON I.InvoiceID = P.InvoiceID WHERE I.Date between @start and @end GROUP BY I.InvoiceID, I.CustomerID, I.Amount, I.Date ORDER BY AmountDue DESC The best equivalent LINQ expression I've come up with, took me much longer to do: var invoices = ( from I in Invoices where I.Date >= start && I.Date <= end join P in Payments on I.InvoiceID equals P.InvoiceID into payments select new{ I.InvoiceID, I.CustomerID, AmountInvoiced = I.Amount, InvoiceDate = I.Date, AmountPaid = ((decimal?)payments.Select(P=>P.Amount).Sum()).GetValueOrDefault(), AmountDue = I.Amount - ((decimal?)payments.Select(P=>P.Amount).Sum()).GetValueOrDefault() } ).OrderByDescending(row=>row.AmountDue); This gets an equivalent result set when run against SQL Server. Using a SQL CE database, however, changes things. The T-SQL stays almost the same. I only have to change ISNULL to COALESCE. Using the same LINQ expression, however, results in an error: There was an error parsing the query. [ Token line number = 4, Token line offset = 9,Token in error = SELECT ] So we look at the generated SQL code: SELECT [t3].[InvoiceID], [t3].[CustomerID], [t3].[Amount] AS [AmountInvoiced], [t3].[Date] AS [InvoiceDate], [t3].[value] AS [AmountPaid], [t3].[value2] AS [AmountDue] FROM ( SELECT [t0].[InvoiceID], [t0].[CustomerID], [t0].[Amount], [t0].[Date], COALESCE(( SELECT SUM([t1].[Amount]) FROM [Payments] AS [t1] WHERE [t0].[InvoiceID] = [t1].[InvoiceID] ),0) AS [value], [t0].[Amount] - (COALESCE(( SELECT SUM([t2].[Amount]) FROM [Payments] AS [t2] WHERE [t0].[InvoiceID] = [t2].[InvoiceID] ),0)) AS [value2] FROM [Invoices] AS [t0] ) AS [t3] WHERE ([t3].[Date] >= @p0) AND ([t3].[Date] <= @p1) ORDER BY [t3].[value2] DESC Ugh! Okay, so it's ugly and inefficient when run against SQL Server, but we're not supposed to care, since it's supposed to be quicker to write, and the performance difference shouldn't be that large. But it just doesn't work against SQL CE, which apparently doesn't support subqueries within the SELECT list. In fact, I've tried several different left join queries in LINQ, and they all seem to have the same problem. Even: from I in Invoices join P in Payments on I.InvoiceID equals P.InvoiceID into payments select new{I, payments} generates: SELECT [t0].[InvoiceID], [t0].[CustomerID], [t0].[Amount], [t0].[Date], [t1].[InvoiceID] AS [InvoiceID2], [t1].[Amount] AS [Amount2], [t1].[Date] AS [Date2], ( SELECT COUNT(*) FROM [Payments] AS [t2] WHERE [t0].[InvoiceID] = [t2].[InvoiceID] ) AS [value] FROM [Invoices] AS [t0] LEFT OUTER JOIN [Payments] AS [t1] ON [t0].[InvoiceID] = [t1].[InvoiceID] ORDER BY [t0].[InvoiceID] which also results in the error: There was an error parsing the query. [ Token line number = 2, Token line offset = 5,Token in error = SELECT ] So how can I do a simple left join on a SQL CE database using LINQ? Am I wasting my time?

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