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• What would you suggest as a high school first language?

  - by ldigas

  Edit by OA: After reading some answers I'll just update the question a little. At first I put it a little bluntly, but some of those gave me some good arguments which have to be taken into consideration while making a stand on this one. (these are mostly picked up from comments and answers below). A few things to take into account: to many pupils this is a first programming language - at this stage most of them have trouble grasping a difference between data types, variable passing, ... and whatnot, less alone pointers and similar 'low level stuff' :) they will all have to pass this to get into next grade (well, big majority of them anyway) not all of them have computers at home, not all of them are willing to learn this, less alone interested in - so the concepts have to be taught on a finite time scale in school hours (as well as practice on computers) free literature is a bonus - the teacher will make some scripts and handaways, but still ... I wouldn't like to bear the parents with the burden of buying expensive literature (also, english is not a native language here ... and although they are all learning it, their ability to read it fluently is somewhat questionable) somebody gave an argument - "a language which does not get in the way of ideas" - good one accessibility on different platforms in not expecially important at this point - although most of the suggested ones are available on win as well as linux - not many macs in this part of europe (their prices are sky high for anything but specialised usage) I will check what are the licencing issues on ms express editions about using it massively in high schools for purposes like this - if someone has any info about this, please, do not be shy with it :) A friend of mine, informatics teacher - in EU it comes as something as junior cs teacher, in a local high school asked me what I thought about what should be the first language pupils should be taught? It is a technical school (a little more oriented towards mathematics than the gymnasium, but not computer oriented totally). So I'm asking you - what do you think should be the first language pupils are exposed to in highschool? They have been teaching Pascal so far, but she's not sure that's a good course. She thought about switching to C (which I resented; considering not all pupils have interests in programming, to start with, and should be taught something higher level since they are just gripping the idea of a loop and such ... for a start), I suggested python or ruby (preferably py since it handles all paradigms). What is your opinion on this one? I looked, but didn't find a similar question on SO, so if there is one, please just point me towards it. Edit: The assumption is that none of the pupils have been exposed to any programming in junior school. See also: What is the best way to teach young kids some basic programming concepts? Best ways to teach a beginner to program How and when do you teach a kid to code What is the easiest language to start with? High School Programming

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• iPhone Image Resources, ICO vs PNG, app bundle filesize

  - by Jasarien

  My application has a collection of around 1940 icons that are used throughout. They're currently in ICO and new images provided to me come in ICO format too. I have noticed that they contain a 16x16 and 32x32 representation of each icon in one file. Each file is roughly 4KB in filesize (as reported by finder, but ls reports that they vary from being ~1000 bytes to 5000 bytes) A very small number of these icons only contain the 32x32 representation, and as a result are only around 700 bytes in size. Currently I am bundling these icons with my application and they are inflating the size of the app a bit more than I would like. Altogether, the images total just about 25.5MB. Xcode must do some kind of compression because the resulting app bundle is about 12.4MB. Compressing this further into a ZIP (as it would be when submitted to the App Store), results in a final file of 5.8MB. I'm aware that the maximum limit for over the air App Store downloads has been raised to 20MB since the introduction of the iPad (I'm not sure if that extends to iPhone apps as well as iPad apps though, if not the limit would be 10MB). My worry is that new icons are going to be added (sometimes up to 10 icons per week), and will continue to inflate the app bundle over time. What is the best way to distribute these icons with my app? Things I've tried and not had much success with: Converting the icons from ICO to PNG: I tried this in the hopes that the pngcrush utility would help out with the filesize. But it appears that it doesn't make much of a difference between a normal PNG and a crushed png (I believe it just optimises the image for display on the iPhone's GPU rather than compress it's size). Also in going from ICO to PNG actually increased the size of the icon file... Zipping the images, and then uncompressing them on first run. While this did reduce the overall image sizes, I found that the effort needed to unzip them, copy them to the documents folder and ensure that duplication doesn't happen on upgrades was too much hassle to be worth the benefit. Also, on original and 3G iPhones unzipping and copying around 25MB of images takes too long and creates a bad experience... Things I've considered but not yet tried: Instead of distributing the icons within the app bundle, host them online, and download each icon on demand (it depends on the user's data as to which icons will actually be displayed and when). Issues with this is that bandwidth costs money, and image downloads will be bandwidth intensive. However, my app currently has a small userbase of around 5,500 users (of which I estimate around 1500 to be active based on Flurry stats), and I have a huge unused bandwidth allowance with my current hosting package. So I'm open to thoughts on how to solve this tricky issue.

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• .NET and C# Exceptions. What is it reasonable to catch.

  - by djna

  Disclaimer, I'm from a Java background. I don't do much C#. There's a great deal of transfer between the two worlds, but of course there are differences and one is in the way Exceptions tend to be thought about. I recently answered a C# question suggesting that under some circstances it's reasonable to do this: try { some work } catch (Exeption e) { commonExceptionHandler(); } (The reasons why are immaterial). I got a response that I don't quite understand: until .NET 4.0, it's very bad to catch Exception. It means you catch various low-level fatal errors and so disguise bugs. It also means that in the event of some kind of corruption that triggers such an exception, any open finally blocks on the stack will be executed, so even if the callExceptionReporter fuunction tries to log and quit, it may not even get to that point (the finally blocks may throw again, or cause more corruption, or delete something important from the disk or database). May I'm more confused than I realise, but I don't agree with some of that. Please would other folks comment. I understand that there are many low level Exceptions we don't want to swallow. My commonExceptionHandler() function could reasonably rethrow those. This seems consistent with this answer to a related question. Which does say "Depending on your context it can be acceptable to use catch(...), providing the exception is re-thrown." So I conclude using catch (Exception ) is not always evil, silently swallowing certain exceptions is. The phrase "Until .NET 4 it is very bad to Catch Exception" What changes in .NET 4? IS this a reference to AggregateException, which may give us some new things to do with exceptions we catch, but I don't think changes the fundamental "don't swallow" rule. The next phrase really bothers be. Can this be right? It also means that in the event of some kind of corruption that triggers such an exception, any open finally blocks on the stack will be executed (the finally blocks may throw again, or cause more corruption, or delete something important from the disk or database) My understanding is that if some low level code had lowLevelMethod() { try { lowestLevelMethod(); } finally { some really important stuff } } and in my code I call lowLevel(); try { lowLevel() } catch (Exception e) { exception handling and maybe rethrowing } Whether or not I catch Exception this has no effect whatever on the excution of the finally block. By the time we leave lowLevelMethod() the finally has already run. If the finally is going to do any of the bad things, such as corrupt my disk, then it will do so. My catching the Exception made no difference. If It reaches my Exception block I need to do the right thing, but I can't be the cause of dmis-executing finallys

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• How do I update with a newly-created detached entity using NHibernate?

  - by Daniel T.

  Explanation: Let's say I have an object graph that's nested several levels deep and each entity has a bi-directional relationship with each other. A -> B -> C -> D -> E Or in other words, A has a collection of B and B has a reference back to A, and B has a collection of C and C has a reference back to B, etc... Now let's say I want to edit some data for an instance ofC. In Winforms, I would use something like this: var instanceOfC; using (var session = SessionFactory.OpenSession()) { // get the instance of C with Id = 3 instanceOfC = session.Linq<C>().Where(x => x.Id == 3); } SendToUIAndLetUserUpdateData(instanceOfC); using (var session = SessionFactory.OpenSession()) { // re-attach the detached entity and update it session.Update(instanceOfC); } In plain English, we grab a persistent instance out of the database, detach it, give it to the UI layer for editing, then re-attach it and save it back to the database. Problem: This works fine for Winform applications because we're using the same entity all throughout, the only difference being that it goes from persistent to detached to persistent again. The problem occurs when I'm using a web service and a browser, sending over JSON data. In this case, the data that comes back is no longer a detached entity, but rather a transient one that just happens to have the same ID as the persistent one. If I use this entity to update, it will wipe out the relationship to B and D unless I sent the entire object graph over to the UI and got it back in one piece. Question: My question is, how do I serialize detached entities over the web, receive them back, and save them, while preserving any relationships that I didn't explicitly change? I know about ISession.SaveOrUpdateCopy and ISession.Merge() (they seem to do the same thing?), but this will still wipe out the relationships if I don't explicitly set them. I could copy the fields from the transient entity to the persistent entity one by one, but this doesn't work too well when it comes to relationships and I'd have to handle version comparisons manually.

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• Matlab cell length

  - by AP

  Ok I seem to have got the most of the problem solved, I just need an expert eye to pick my error as I am stuck. I have a file of length [125 X 27] and I want to convert it to a file of length [144 x 27]. Now, I want to replace the missing files (time stamps) rows of zeros. (ideally its a 10 min daily average thus should have file length of 144) Here is the code I am using: fid = fopen('test.csv', 'rt'); data = textscan(fid, ['%s' repmat('%f',1,27)], 'HeaderLines', 1, 'Delimiter', ','); fclose(fid); %//Make time a datenum of the first column time = datenum(data{1} , 'mm/dd/yyyy HH:MM') %//Find the difference in minutes from each row timeDiff = round(diff(datenum(time)*(24*60))) %//the rest of the data data = cell2mat(data(2:28)); newdata=zeros(144,27); for n=1:length(timeDiff) if timeDiff(n)==10 newdata(n,:)=data(n,:); newdata(n+1,:)=data(n+1,:); else p=timeDiff(n)/10 n=n+p; end end Can somebody please help me to find the error inside my for loop. My output file seems to miss few timestamped values. %*********************************************************************************************************** Can somebody help me to figure out the uiget to read the above file?? i am replacing fid = fopen('test.csv', 'rt'); data = textscan(fid, ['%s' repmat('%f',1,27)], 'HeaderLines', 1, 'Delimiter', ','); fclose(fid); With [c,pathc]=uigetfile({'*.txt'},'Select the file','C:\data'); file=[pathc c]; file= textscan(c, ['%s' repmat('%f',1,27)], 'HeaderLines', 1, 'Delimiter', ','); And its not working % NEW ADDITION to old question p = 1; %index into destination for n = 1:length(timeDiff) % if timeDiff(n) == 10 % newfile(p,:) = file(n,:); % newfile(p+1,:)=file(n+1,:); % p = p + 1; % else % p = p + (timeDiff(n)/10); % end q=cumsum(timeDiff(n)/10); if q==1 newfile(p,:)=file(n,:); p=p+1; else p = p + (timeDiff(n)/10); end end xlswrite('testnewws11.xls',newfile); even with the cumsum command this code fails when my file has 1,2 time stamps in middle of long missing ones example 8/16/2009 0:00 5.34 8/16/2009 0:10 3.23 8/16/2009 0:20 2.23 8/16/2009 0:30 1.23 8/16/2009 0:50 70 8/16/2009 2:00 5.23 8/16/2009 2:20 544 8/16/2009 2:30 42.23 8/16/2009 3:00 71.23 8/16/2009 3:10 3.23 My output looks like 5.34 3.23 2.23 0 0 0 0 0 0 0 0 0 5.23 544. 42.23 0 0 0 3.23 Any ideas?

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• How do I add Objective C code to a FireBreath Project?

  - by jmort253

  I am writing a browser plugin for Mac OS that will place a status bar icon in the status bar, which users can use to interface with the browser plugin. I've successfully built a FireBreath 1.6 project in XCode 4.4.1, and can install it in the browser. However, FireBreath uses C++, whereas a large majority of the existing libraries for Mac OS are written in Objective C. In the /Mac/projectDef.make file, I added the Cocoa Framework and Foundation Framework, as suggested here and in other resources I've found on the Internet: target_link_libraries(${PROJECT_NAME} ${PLUGIN_INTERNAL_DEPS} ${Cocoa.framework} # added line ${Foundation.framework} # added line ) I reran prepmac.sh, expecting a new project to be created in XCode with my .mm files, and .m files; however, it seems that they're being ignored. I only see the .cpp and .h files. I added rules for those in the projectDef.make file, but it doesn't seem to make a difference: file (GLOB PLATFORM RELATIVE ${CMAKE_CURRENT_SOURCE_DIR} Mac/[^.]*.cpp Mac/[^.]*.h Mac/[^.]*.m #added by me Mac/[^.]*.mm #added by me Mac/[^.]*.cmake ) Even if I add the files in manually, I get a series of compilation errors. There are about 20 of them, all related to the file NSObjRuntime.h file: Parse Issue - Expected unqualified-id Parse Issue - Unknown type name 'NSString' Semantic Issue - Use of undeclared identifier 'NSString' Parse Issue - Unknown type name 'NSString' ... ... Semantic Issue - Use of undeclared identifier 'aSelectorName' ... ... Semantic Issue - Use of undeclared identifier 'aClassName' ... It continues like this for some time with similar errors... From what I've read, these errors appear because of dependencies on the Foundation Framework, which I believe I've included in the project. I also tried clicking the project in XCode I'm to the point now where I'm not sure what to try next. People say it's not hard to use Objective C in C/C++ code, but being new to XCode and Objective C might contribute to my confusion. This is only day 4 for me in this new platform. What do I need to do to get XCode to compile the Objective C code? Please remember that I'm a little new to this, so I'd appreciate it if you leave detailed answers as opposed to the vague one-liners that are common in the firebreath tag. I'm just a little in over my head, but if you can get me past this hurdle I'm certain I'll be good to go from there. UPDATE: I edited projects/MyPlugin/CMakeLists.txt and added in the .m and .mm rules there too. after running prepmac.sh, the files are included in the project, but I still get the same compile errors. I moved all the .h files and .mm files from the Obj C code to the MyPlugin root folder and reran the prepmac.sh file. Problem still exists. Same compile errors.

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• Grid sorting with persistent master sort

  - by MikeWyatt

  I have a UI with a grid. Each record in the grid is sorted by a "master" sort column, let's call it a page number. Each record is a story in a magazine. I want the user to be able to drag and drop a record to a new position in the grid and automatically update the page number field to reflect the updated position. Easy enough, right? Now imagine that I also want to have the grid sortable by any other column (story title, section, author name, etc.). How does the drag and drop operation work now? Revert to page number sort during or after the drag and drop operation? This could confuse the user (why did my sort just change?). It would also result in arbitrary row positioning. Would the story now be before the row that was after it when the user dropped it? Or, would it be after the row that was before it? Those rows may now be widely separated after the master order sort. Disable the drag and drop feature if the grid isn't currently sorted by the page number? This would be easy, but the user might wonder why he can't drag and drop at certain times. Knowing to first sort by page number may not be very intuitive. Let the user rearrange his rows, but not make any changes to the page number? Require the user to enter a "Arrange Stories" mode, in which the grid sort is temporarily switched to page number and drag and drop is enabled? They would then exit the mode, and the previous sort would be reapplied. The big difference between this and the second option is that it would be more explicit than simply clicking on a column header. Any other ideas, or reasons why one of the above is the way to go? EDIT I'd like to point out that any of the above is technically possible, and easy to implement. My question is design-related. What is the most intuitive way to solve this problem, from the user's perspective?

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• Greasemonkey is getting an empty document.body on select Google pages.

  - by Brock Adams

  Hi, I have a Greasemonkey script that processes Google search results. But it's failing in a few instances, when xpath searches (and document body) appear to be empty. Running the code in Firebug's console works every time. It only fails in a Greasemonkey script. Greasemonkey sees an empty document.body. I've boiled the problem down to a test, greasemonkey script, below. I'm using Firefox 3.5.9 and Greasemonkey 0.8.20100408.6 (but earlier versions had the same problem). Problem: Greasemonkey sees an empty document.body. Recipe to Duplicate: Install the Greasemonkey script. Open a new tab or window. Navigate to Google.com (http://www.google.com/). Search on a simple term like "cats". Check Firefox's Error console (Ctrl-shift-J) or Firebug's console. The script will report that document body is empty. Hit refresh. The script will show a good result (document body found). Note that the failure only reliably appears on Google results obtained this way, and on a new tab/window. Turn javascript off globally (javascript.enabled set to false in about:config). Repeat steps 2 thru 5. Only now the Greasemonkey script will work. It seems that Google javascript is killing the DOM tree for greasemonkey, somehow. I've tried a time-delayed retest and even a programmatic refresh; the script still fails to see the document body. Test Script: // // ==UserScript== // @name TROUBLESHOOTING 2 snippets // @namespace http://www.google.com/ // @description For code that has funky misfires and defies standard debugging. // @include http://*/* // ==/UserScript== // function LocalMain (sTitle) { var sUserMessage = ''; //var sRawHtml = unsafeWindow.document.body.innerHTML; //-- unsafeWindow makes no difference. var sRawHtml = document.body.innerHTML; if (sRawHtml) { sRawHtml = sRawHtml.replace (/^\s\s*/, ''). substr (0, 60); sUserMessage = sTitle + ', Doc body = ' + sRawHtml + ' ...'; } else { sUserMessage = sTitle + ', Document body seems empty!'; } if (typeof (console) != "undefined") { console.log (sUserMessage); } else { if (typeof (GM_log) != "undefined") GM_log (sUserMessage); else if (!sRawHtml) alert (sUserMessage); } } LocalMain ('Preload'); window.addEventListener ("load", function() {LocalMain ('After load');}, false);

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• How do you return a pointer to a base class with a virtual function?

  - by Nick Sweet

  I have a base class called Element, a derived class called Vector, and I'm trying to redefine two virtual functions from Element in Vector. //element.h template <class T> class Element { public: Element(); virtual Element& plus(const Element&); virtual Element& minus(const Element&); }; and in another file //Vector.h #include "Element.h" template <class T> class Vector: public Element<T> { T x, y, z; public: //constructors Vector(); Vector(const T& x, const T& y = 0, const T& z =0); Vector(const Vector& u); ... //operations Element<T>& plus(const Element<T>& v) const; Element<T>& minus(const Element<T>& v) const; ... }; //sum template <class T> Element<T>& Vector<T>::plus(const Element<T>& v) const { Element<T>* ret = new Vector((x + v.x), (y + v.y), (z + v.z)); return *ret; } //difference template <class T> Element<T>& Vector<T>::minus(const Element<T>& v) const { Vector<T>* ret = new Vector((x - v.x), (y - v.y), (z - v.z)); return *ret; } but I always get error: 'const class Element' has no member named 'getx' So, can I define my virtual functions to take Vector& as an argument instead, or is there a way for me to access the data members of Vector through a pointer to Element? I'm still fairly new to inheritance polymorphism, fyi.

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• How do compare dates when one of those are in string format in android

  - by Raj

  I am very much new to android so need some good help with a code example. I am getting a date in form of string from a server in the following format 2012-08-17 00:00:00 I want to compare this string with current date to find the difference between the dates in the form of year, months and days... I tried playing around it in the following code Date currentDate = new Date(System.currentTimeMillis()); Log.v("@@@@@@@@@","Current Date: " + currentDate); Date passDate = new SimpleDateFormat().parse(passDateString); Log.v("@@@@@@@@@","Pass Date: " + passDate); dateDifference = passDate.compareTo(currentDate); but it returned with following exception 04-15 12:08:29.101: V/@@@@@@@@@(1161): Current Date: Sun Apr 15 12:08:29 GMT+01:00 2012 04-15 12:08:29.101: W/System.err(1161): java.text.ParseException: Unparseable date: 2012-08-17 00:00:00 04-15 12:08:29.111: W/System.err(1161): at java.text.DateFormat.parse(DateFormat.java:645) 04-15 12:08:29.111: W/System.err(1161): at org.apis.PassesListItemAdapter.getView(PassesListItemAdapter.java:77) 04-15 12:08:29.111: W/System.err(1161): at android.widget.AbsListView.obtainView(AbsListView.java:1315) 04-15 12:08:29.111: W/System.err(1161): at android.widget.ListView.makeAndAddView(ListView.java:1727) 04-15 12:08:29.111: W/System.err(1161): at android.widget.ListView.fillDown(ListView.java:652) 04-15 12:08:29.111: W/System.err(1161): at android.widget.ListView.fillFromTop(ListView.java:709) 04-15 12:08:29.111: W/System.err(1161): at android.widget.ListView.layoutChildren(ListView.java:1580) 04-15 12:08:29.111: W/System.err(1161): at android.widget.AbsListView.onLayout(AbsListView.java:1147) 04-15 12:08:29.111: W/System.err(1161): at android.view.View.layout(View.java:7034) 04-15 12:08:29.111: W/System.err(1161): at android.widget.RelativeLayout.onLayout(RelativeLayout.java:909) 04-15 12:08:29.111: W/System.err(1161): at android.view.View.layout(View.java:7034) 04-15 12:08:29.111: W/System.err(1161): at android.widget.FrameLayout.onLayout(FrameLayout.java:333) 04-15 12:08:29.111: W/System.err(1161): at android.view.View.layout(View.java:7034) 04-15 12:08:29.111: W/System.err(1161): at android.widget.FrameLayout.onLayout(FrameLayout.java:333) 04-15 12:08:29.111: W/System.err(1161): at android.view.View.layout(View.java:7034) 04-15 12:08:29.111: W/System.err(1161): at android.view.ViewRoot.performTraversals(ViewRoot.java:1049) 04-15 12:08:29.111: W/System.err(1161): at android.view.ViewRoot.handleMessage(ViewRoot.java:1744) 04-15 12:08:29.111: W/System.err(1161): at android.os.Handler.dispatchMessage(Handler.java:99) 04-15 12:08:29.111: W/System.err(1161): at android.os.Looper.loop(Looper.java:144) 04-15 12:08:29.111: W/System.err(1161): at android.app.ActivityThread.main(ActivityThread.java:4937) 04-15 12:08:29.111: W/System.err(1161): at java.lang.reflect.Method.invokeNative(Native Method) 04-15 12:08:29.111: W/System.err(1161): at java.lang.reflect.Method.invoke(Method.java:521) 04-15 12:08:29.111: W/System.err(1161): at com.android.internal.os.ZygoteInit$MethodAndArgsCaller.run(ZygoteInit.java:868) 04-15 12:08:29.111: W/System.err(1161): at com.android.internal.os.ZygoteInit.main(ZygoteInit.java:626) 04-15 12:08:29.111: W/System.err(1161): at dalvik.system.NativeStart.main(Native Method) I am stuck... please help Raj

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• GCC problem with raw double type comparisons

  - by Monomer

  I have the following bit of code, however when compiling it with GCC 4.4 with various optimization flags I get some unexpected results when its run. #include <iostream> int main() { const unsigned int cnt = 10; double lst[cnt] = { 0.0 }; const double v[4] = { 131.313, 737.373, 979.797, 731.137 }; for(unsigned int i = 0; i < cnt; ++i) { lst[i] = v[i % 4] * i; } for(unsigned int i = 0; i < cnt; ++i) { double d = v[i % 4] * i; if(lst[i] != d) { std::cout << "error @ : " << i << std::endl; return 1; } } return 0; } when compiled with: "g++ -pedantic -Wall -Werror -O1 -o test test.cpp" I get the following output: "error @ : 3" when compiled with: "g++ -pedantic -Wall -Werror -O2 -o test test.cpp" I get the following output: "error @ : 3" when compiled with: "g++ -pedantic -Wall -Werror -O3 -o test test.cpp" I get no errors when compiled with: "g++ -pedantic -Wall -Werror -o test test.cpp" I get no errors I do not believe this to be an issue related to rounding, or epsilon difference in the comparison. I've tried this with Intel v10 and MSVC 9.0 and they all seem to work as expected. I believe this should be nothing more than a bitwise compare. If I replace the if-statement with the following: if (static_cast<long long int>(lst[i]) != static_cast<long long int>(d)), and add "-Wno-long-long" I get no errors in any of the optimization modes when run. If I add std::cout << d << std::endl; before the "return 1", I get no errors in any of the optimization modes when run. Is this a bug in my code, or is there something wrong with GCC and the way it handles the double type?

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• How to cache queries in EJB and return result efficient (performance POV)

  - by Maxym

  I use JBoss EJB 3.0 implementation (JBoss 4.2.3 server) At the beginning I created native query all the time using construction like Query query = entityManager.createNativeQuery("select * from _table_"); Of couse it is not that efficient, I performed some tests and found out that it really takes a lot of time... Then I found a better way to deal with it, to use annotation to define native queries: @NamedNativeQuery( name = "fetchData", value = "select * from _table_", resultClass=Entity.class ) and then just use it Query query = entityManager.createNamedQuery("fetchData"); the performance of code line above is two times better than where I started from, but still not that good as I expected... then I found that I can switch to Hibernate annotation for NamedNativeQuery (anyway, JBoss's implementation of EJB is based on Hibernate), and add one more thing: @NamedNativeQuery( name = "fetchData2", value = "select * from _table_", resultClass=Entity.class, readOnly=true) readOnly - marks whether the results are fetched in read-only mode or not. It sounds good, because at least in this case of mine I don't need to update data, I wanna just fetch it for report. When I started server to measure performance I noticed that query without readOnly=true (by default it is false) returns result with each iteration better and better, and at the same time another one (fetchData2) works like "stable" and with time difference between them is shorter and shorter, and after 5 iterations speed of both was almost the same... The questions are: 1) is there any other way to speed query using up? Seems that named queries should be prepared once, but I can't say it... In fact if to create query once and then just use it it would be better from performance point of view, but it is problematic to cache this object, because after creating query I can set parameters (when I use ":variable" in query), and it changes query object (isn't it?). well, is here any way to cache them? Or named query is the best option I can use? 2) any other approaches how to make results retrieveng faster. I mean, for instance I don't need those Entities to be attached, I won't update them, all I need is just fetch collection of data. Maybe readOnly is the only available way, so I can't speed it up, but who knows :) P.S. I don't ask about DB performance, all I need now is how not to create query all the time, so use it efficient, and to "allow" EJB to do less job with the same result concerning data returning.

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• Any thoughts on how to create a true 'punch-out' area in a Sprite?

  - by rhtx

  I've been working on this for awhile, now. You might also call it a 'reverse mask', or an 'inverse mask'. Basically, I'm creating a view window within a display object. I need to allow objects on the stage that are under the window to be able to interact with the mouse. This is similar to a WPF question: http://stackoverflow.com/questions/740994/use-wpf-object-to-punch-hole-in-another, which has a much shorter write-up. I've got a Class called PunchOutShield, which creates a Sprite that covers the stage (or over some desired area). The Sprite's Graphics object is filled using the color and transparency of Flex's modal screen. The result is a screen that looks like the screen which appears behind a modal PopUp. PunchOutShield has a method called punch, which takes two arguments - the first is a Shape object, which defines the shape of the punch-through area; the second is a Point object, which indicates where to position the punch-through area. It took some experimenting, but I found that I can successfully create a punch-out area (i.e. - the modal screen does not display within the bounds of the given Shape). To do this, I set cacheAsBitmap to true on the Sprite that is used to create the modal screen, and also on the Shape object, which is added to the modal screen Sprite's displayList. If I set the blend mode of the Shape to ERASE, a completely transparent area is created in the modal screen. So far, great. The problem is that Shape does not subclass InteractiveObject, so there is no way to set mouseEnabled = false on it. And so, it prevents interaction between the mouse and any objects that are visible through the punch-out area. On top of that, InteractiveObject isn't available to look at, so I can't see if there is a way to borrow what it's doing to provide the mouseEnabled functionality and apply it to a subclass of Shape. I've tried using another Sprite object, rather than a Shape object, but the blending doesn't work out correctly. I'm not sure why there is a difference, but the Shape object seems to somehow combine with the parenting Sprite, allowing the ERASE blendMode to effect the desired punch-out visual appearance. It wouldn't be the end of the world if I had to draw up the screen with a series of rectangles so that the punch-out area was just simply not drawn, but that approach won't work if the punch-out area is complex. Or round. Any thoughts on this approach, or on an alternative approach?

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• Calling a function that resides in the main page from a plugin?

  - by Justin Lee

  I want to call a function from within plugin, but the function is on the main page and not the plugin's .js file. EDIT I have jQuery parsing a very large XML file and building, subsequently, a large list (1.1 MB HTML file when dynamic content is copied, pasted, then saved) that has expand/collapse functionality through a plugin. The overall performance on IE is super slow and doggy, assuming since the page/DOM is so big. I am currently trying to save the collapsed content in the event.data when it is collapsed and remove it from the DOM, then bring it back when it is told to expand... the issue that I am having is that when I bring the content back, obviously the "click" and "hover" events are gone. I'm trying to re-assign them, currently doing so inside the plugin after the plugin expands the content. The issue then though is that is says the function that I declare within the .click() is not defined. Also the hover event doesn't seem to be re-assigning either.... if ($(event.data.trigger).attr('class').indexOf('collapsed') != -1 ) { // if expanding // console.log(event.data.targetContent); $(event.data.trigger).after(event.data.targetContent); $(event.data.target).hide(); /* This Line --->*/ $(event.data.target + 'a.addButton').click(addResourceToList); $(event.data.target + 'li.resource') .hover( function() { if (!($(this).attr("disabled"))) { $(this).addClass("over"); $(this).find("a").css({'display':'block'}); } }, function () { if (!($(this).attr("disabled"))) { $(this).removeClass("over"); $(this).children("a").css({'display':'none'}); } } ); $(event.data.target).css({ "height": "0px", "padding-top": "0px", "padding-bottom": "0px", "margin-top": "0px", "margin-bottom": "0px"}); $(event.data.target).show(); $(event.data.target).animate({ height: event.data.heightVal + "px", paddingTop: event.data.topPaddingVal + "px", paddingBottom: event.data.bottomPaddingVal + "px", marginTop: event.data.topMarginVal + "px", marginBottom: event.data.bottomMarginVal + "px"}, "normal");//, function(){$(this).hide();}); $(event.data.trigger).removeClass("collapsed"); $.cookies.set('jcollapserSub_' + event.data.target, 'expanded', {hoursToLive: 24 * 365}); } else if ($(event.data.trigger).attr('class').indexOf('collapsed') == -1 ) { // if collapsing $(event.data.target).animate({ height: "0px", paddingTop: "0px", paddingBottom: "0px", marginTop: "0px", marginBottom: "0px"}, "normal", function(){$(this).hide();$(this).remove();}); $(event.data.trigger).addClass("collapsed"); $.cookies.set('jcollapserSub_' + event.data.target, 'collapsed', {hoursToLive: 24 * 365}); } EDIT So, having new eyes truly makes a difference. As I was reviewing the code in this post this morning after being away over the weekend, I found where I had err'd. This: $(event.data.target + 'a.addButton').click(addResourceToList); Should be this (notice the space before a.addbutton): $(event.data.target + ' a.addButton').click(addResourceToList); Same issue with the "li.resource". So it was never pointing to the right elements... Thank you, Rene, for your help!!

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• Feedback on Optimizing C# NET Code Block

  - by Brett Powell

  I just spent quite a few hours reading up on TCP servers and my desired protocol I was trying to implement, and finally got everything working great. I noticed the code looks like absolute bollocks (is the the correct usage? Im not a brit) and would like some feedback on optimizing it, mostly for reuse and readability. The packet formats are always int, int, int, string, string. try { BinaryReader reader = new BinaryReader(clientStream); int packetsize = reader.ReadInt32(); int requestid = reader.ReadInt32(); int serverdata = reader.ReadInt32(); Console.WriteLine("Packet Size: {0} RequestID: {1} ServerData: {2}", packetsize, requestid, serverdata); List<byte> str = new List<byte>(); byte nextByte = reader.ReadByte(); while (nextByte != 0) { str.Add(nextByte); nextByte = reader.ReadByte(); } // Password Sent to be Authenticated string string1 = Encoding.UTF8.GetString(str.ToArray()); str.Clear(); nextByte = reader.ReadByte(); while (nextByte != 0) { str.Add(nextByte); nextByte = reader.ReadByte(); } // NULL string string string2 = Encoding.UTF8.GetString(str.ToArray()); Console.WriteLine("String1: {0} String2: {1}", string1, string2); // Reply to Authentication Request MemoryStream stream = new MemoryStream(); BinaryWriter writer = new BinaryWriter(stream); writer.Write((int)(1)); // Packet Size writer.Write((int)(requestid)); // Mirror RequestID if Authenticated, -1 if Failed byte[] buffer = stream.ToArray(); clientStream.Write(buffer, 0, buffer.Length); clientStream.Flush(); } I am going to be dealing with other packet types as well that are formatted the same (int/int/int/str/str), but different values. I could probably create a packet class, but this is a bit outside my scope of knowledge for how to apply it to this scenario. If it makes any difference, this is the Protocol I am implementing. http://developer.valvesoftware.com/wiki/Source_RCON_Protocol

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• Linux configurations that would affect Java memory usage?

  - by wmacura

  Hi, Background: I have a set of java background workers I start as part of my webapp. I develop locally on Ubuntu 10.10 and deploy to an Ubuntu 10.04LTS server (a media temple (ve) instance). They're both running the same JVM: Sun JVM 1.6.0_22-b04. As part of the initialization script each worker is started with explicit Xmx, Xms, and XX:MaxPermGen settings. Yet somehow locally all 10 workers use 250MB, while on the server they use more than 2.7GB. I don't know how to begin to track this down. I thought the Ubuntu (and thus, kernel) version might make a difference, but I tried an old 10.04 VM and it behaves as expected. I've noticed that the machine does not seem to ever use memory for buffer or cache (according to htop), which seems a bit strange, but perhaps normal for a server? (edited) Some info: (server) root@devel:/app/axir/target# uname -a Linux devel 2.6.18-028stab069.5 #1 SMP Tue May 18 17:26:16 MSD 2010 x86_64 GNU/Linux (local) wiktor@beastie:~$ uname -a Linux beastie 2.6.35-25-generic #44-Ubuntu SMP Fri Jan 21 17:40:44 UTC 2011 x86_64 GNU/Linux (edited) Comparing PS output: (ps -eo "ppid,pid,cmd,rss,sz,vsz") PPID PID CMD RSS SZ VSZ (local) 1588 1615 java -cp axir-distribution. 25484 234382 937528 1615 1631 java -cp /home/wiktor/Code/ 83472 163059 652236 1615 1657 java -cp /home/wiktor/Code/ 70624 89135 356540 1615 1658 java -cp /home/wiktor/Code/ 37652 77625 310500 1615 1669 java -cp /home/wiktor/Code/ 38096 77733 310932 1615 1675 java -cp /home/wiktor/Code/ 37420 61395 245580 1615 1684 java -cp /home/wiktor/Code/ 38000 77736 310944 1615 1703 java -cp /home/wiktor/Code/ 39180 78060 312240 1615 1712 java -cp /home/wiktor/Code/ 38488 93882 375528 1615 1719 java -cp /home/wiktor/Code/ 38312 77874 311496 1615 1726 java -cp /home/wiktor/Code/ 38656 77958 311832 1615 1727 java -cp /home/wiktor/Code/ 78016 89429 357716 (server) 22522 23560 java -cp axir-distribution. 24860 285196 1140784 23560 23585 java -cp /app/axir/target/a 100764 161629 646516 23560 23667 java -cp /app/axir/target/a 72408 92682 370728 23560 23670 java -cp /app/axir/target/a 39948 97671 390684 23560 23674 java -cp /app/axir/target/a 40140 81586 326344 23560 23739 java -cp /app/axir/target/a 39688 81542 326168 They look very similar. In fact, the question now is why, if I add up the virtual memory usage on the server (3.2GB) does it more closely reflect 2.4GB of memory used (according to free), yet locally the virtual memory used adds up to a much more substantial 4.7GB but only actually uses ~250MB. It seems that perhaps memory isn't being shared as aggressively. (if that's even possible) Thank you for your help, Wiktor

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• Stored procedure performance randomly plummets; trivial ALTER fixes it. Why?

  - by gWiz

  I have a couple of stored procedures on SQL Server 2005 that I've noticed will suddenly take a significantly long time to complete when invoked from my ASP.NET MVC app running in an IIS6 web farm of four servers. Normal, expected completion time is less than a second; unexpected anomalous completion time is 25-45 seconds. The problem doesn't seem to ever correct itself. However, if I ALTER the stored procedure (even if I don't change anything in the procedure, except to perhaps add a space to the script created by SSMS Modify command), the completion time reverts to expected completion time. IIS and SQL Server are running on separate boxes, both running Windows Server 2003 R2 Enterprise Edition. SQL Server is Standard Edition. All machines have dual Xeon E5450 3GHz CPUs and 4GB RAM. SQL Server is accessed using its TCP/IP protocol over gigabit ethernet (not sure what physical medium). The problem is present from all web servers in the web farm. When I invoke the procedure from a query window in SSMS on my development machine, the procedure completes in normal time. This is strange because I was under the impression that SSMS used the same SqlClient driver as in .NET. When I point my development instance of the web app to the production database, I again get the anomalous long completion time. If my SqlCommand Timeout is too short, I get System.Data.SqlClient.SqlException: Timeout expired. The timeout period elapsed prior to completion of the operation or the server is not responding. Question: Why would performing ALTER on the stored procedure, without actually changing anything in it, restore the completion time to less than a second, as expected? Edit: To clarify, when the procedure is running slow for the app, it simultaneously runs fine in SSMS with the same parameters. The only difference I can discern is login credentials (next time I notice the behavior, I'll be checking from SSMS with the same creds). The ultimate goal is to get the procs to sustainably run with expected speed without requiring occasional intervention. Resolution: I wanted to to update this question in case others are experiencing this issue. Following the leads of the answers below, I was able to consistently reproduce this behavior. In order to test, I utilize sp_recompile and pass it one of the susceptible sprocs. I then initiate a website request from my browser that will invoke the sproc with atypical parameters. Lastly, I initiate a website request to a page that invokes the sproc with typical parameters, and observe that the request does not complete because of a SQL timeout on the sproc invocation. To resolve this on SQL Server 2005, I've added OPTIMIZE FOR hints to my SELECT. The sprocs that were vulnerable all have the "all-in-one" pattern described in this article. This pattern is certainly not ideal but was a necessary trade-off given the timeframe for the project.

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• How to download file into string with progress callback?

  - by Kaminari

  I would like to use the WebClient (or there is another better option?) but there is a problem. I understand that opening up the stream takes some time and this can not be avoided. However, reading it takes a strangely much more amount of time compared to read it entirely immediately. Is there a best way to do this? I mean two ways, to string and to file. Progress is my own delegate and it's working good. FIFTH UPDATE: Finally, I managed to do it. In the meantime I checked out some solutions what made me realize that the problem lies elsewhere. I've tested custom WebResponse and WebRequest objects, library libCURL.NET and even Sockets. The difference in time was gzip compression. Compressed stream lenght was simply half the normal stream lenght and thus download time was less than 3 seconds with the browser. I put some code if someone will want to know how i solved this: (some headers are not needed) public static string DownloadString(string URL) { WebClient client = new WebClient(); client.Headers["User-Agent"] = "Mozilla/5.0 (Windows; U; Windows NT 6.1; en-US) AppleWebKit/532.5 (KHTML, like Gecko) Chrome/4.1.249.1045 Safari/532.5"; client.Headers["Accept"] = "application/xml,application/xhtml+xml,text/html;q=0.9,text/plain;q=0.8,image/png,*/*;q=0.5"; client.Headers["Accept-Encoding"] = "gzip,deflate,sdch"; client.Headers["Accept-Charset"] = "ISO-8859-2,utf-8;q=0.7,*;q=0.3"; Stream inputStream = client.OpenRead(new Uri(URL)); MemoryStream memoryStream = new MemoryStream(); const int size = 32 * 4096; byte[] buffer = new byte[size]; if (client.ResponseHeaders["Content-Encoding"] == "gzip") { inputStream = new GZipStream(inputStream, CompressionMode.Decompress); } int count = 0; do { count = inputStream.Read(buffer, 0, size); if (count > 0) { memoryStream.Write(buffer, 0, count); } } while (count > 0); string result = Encoding.Default.GetString(memoryStream.ToArray()); memoryStream.Close(); inputStream.Close(); return result; } I think that asyncro functions will be almost the same. But i will simply use another thread to fire this function. I dont need percise progress indication.

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• Can I write a test that succeeds if and only if a statement does not compile?

  - by Billy ONeal

  I'd like to prevent clients of my class from doing something stupid. To that end, I have used the type system, and made my class only accept specific types as input. Consider the following example (Not real code, I've left off things like virtual destructors for the sake of example): class MyDataChunk { //Look Ma! Implementation! }; class Sink; class Source { virtual void Run() = 0; Sink *next_; void SetNext(Sink *next) { next_ = next; } }; class Sink { virtual void GiveMeAChunk(const MyDataChunk& data) { //Impl }; }; class In { virtual void Run { //Impl } }; class Out { }; //Note how filter and sorter have the same declaration. Concrete classes //will inherit from them. The seperate names are there to ensure only //that some idiot doesn't go in and put in a filter where someone expects //a sorter, etc. class Filter : public Source, public Sink { //Drop objects from the chain-of-command pattern that don't match a particular //criterion. }; class Sorter : public Source, public Sink { //Sorts inputs to outputs. There are different sorters because someone might //want to sort by filename, size, date, etc... }; class MyClass { In i; Out o; Filter f; Sorter s; public: //Functions to set i, o, f, and s void Execute() { i.SetNext(f); f.SetNext(s); s.SetNext(o); i.Run(); } }; What I don't want is for somebody to come back later and go, "Hey, look! Sorter and Filter have the same signature. I can make a common one that does both!", thus breaking the semantic difference MyClass requires. Is this a common kind of requirement, and if so, how might I implement a test for it?

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• Best way to version control a WCF application with Git?

  - by Sam

  Suppose I have the following projects. The format is [ProjectName] : [ProjectDependency1, ProjectDependency2, etc.] // Service CoolLibrary WcfApp.Core WcfApp.Contracts WcfApp.Services : CoolLibrary, WcfApp.Core, WcfApp.Contracts // Clients CustomerX.App : WcfApp.Contracts CustomerY.App : WcfApp.Contracts CustomerZ.App : WcfApp.Contracts (On a side note, WcfApp.Contracts should not depend on WcfApp.Core, right? Else CustomerX.App would also depend on and thus be exposed to the service domain model?) (CoolLibrary is shared with other applications, so I can't just put it inside of WcfApp.Services.) All of this code is in-house. I was thinking of having 6 repositories for this. The format is [repository folder name] : [Projects included in repository.] 1. CoolLibrary.git : CoolLibrary 2. WcfApp.Contracts.git : WcfApp.Contracts 3. WcfApp.git : WcfApp.Core, WcfApp.Services 4. CustomerX.App.git : CustomerX.App 5. CustomerY.App.git : CustomerY.App 6. CustomerZ.App.git : CustomerZ.App How should I manage my project dependencies? I see three options: I could use binaries which I have to manually copy to each dependent repository. This would be easiest at the start, but my repositories would be a little bloated, and it'd become more tedious as I add more client apps for customers. I could import dependent code as submodules. This is what I will probably end up doing, although I keep reading on the web that submodules are a hassle. I also read that I can use something called the subtree merge strategy, but I am not sure how it is different from just cloning the repo into a subdirectory and adding the subdirectory to .gitignore. Is the difference that the subtree is recorded in the master repository, so (for example) cloning it from a different location will also pull the subtree? I know I asked a lot of questions in this post, but the most important two questions I have are: 1. Am I using the right number and layout of repositories? Should I use less or more? 2. Which of the three dependency management strategies would you recommend? Is there another strategy I haven't considered?

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• CSS selectors : should I minimise my use of the class attribute in the HTML or optimise the speed

  - by Laurent Bourgault-Roy

  As I was working on a small website, I decided to use the PageSpeed extension to check if their was some improvement I could do to make the site load faster. However I was quite surprise when it told me that my use of CSS selector was "inefficient". I was always told that you should keep the usage of the class attribute in the HTML to a minimum, but if I understand correctly what PageSpeed tell me, it's much more efficient for the browser to match directly against a class name. It make sense to me, but it also mean that I need to put more CSS classes in my HTML. It also make my .css file a little harder to read. I usually tend to mark my CSS like this : #mainContent p.productDescription em.priceTag { ... } Which make it easy to read : I know this will affect the main content and that it affect something in a paragraph tag (so I wont start to put all sort of layout code in it) that describe a product and its something that need emphasis. However it seem I should rewrite it as .priceTag { ... } Which remove all context information about the style. And if I want to use differently formatted price tag (for example, one in a list on the sidebar and one in a paragraph), I need to use something like that .paragraphPriceTag { ... } .listPriceTag { ... } Which really annoy me since I seem to duplicate the semantic of the HTML in my classes. And that mean I can't put common style in an unqualified .priceTag { ... } and thus I need to replicate the style in both CSS rule, making it harder to make change. (Altough for that I could use multiple class selector, but IE6 dont support them) I believe making code harder to read for the sake of speed has never been really considered a very good practice . Except where it is critical, of course. This is why people use PHP/Ruby/C# etc. instead of C/assembly to code their site. It's easier to write and debug. So I was wondering if I should stick with few CSS classes and complex selector or if I should go the optimisation route and remove my fancy CSS selectors for the sake of speed? Does PageSpeed make over the top recommandation? On most modern computer, will it even make a difference?

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• Simplifying const Overloading?

  - by templatetypedef

  Hello all- I've been teaching a C++ programming class for many years now and one of the trickiest things to explain to students is const overloading. I commonly use the example of a vector-like class and its operator[] function: template <typename T> class Vector { public: T& operator[] (size_t index); const T& operator[] (size_t index) const; }; I have little to no trouble explaining why it is that two versions of the operator[] function are needed, but in trying to explain how to unify the two implementations together I often find myself wasting a lot of time with language arcana. The problem is that the only good, reliable way that I know how to implement one of these functions in terms of the other is with the const_cast/static_cast trick: template <typename T> const T& Vector<T>::operator[] (size_t index) const { /* ... your implementation here ... */ } template <typename T> T& Vector<T>::operator[] (size_t index) { return const_cast<T&>(static_cast<const Vector&>(*this)[index]); } The problem with this setup is that it's extremely tricky to explain and not at all intuitively obvious. When you explain it as "cast to const, then call the const version, then strip off constness" it's a little easier to understand, but the actual syntax is frightening,. Explaining what const_cast is, why it's appropriate here, and why it's almost universally inappropriate elsewhere usually takes me five to ten minutes of lecture time, and making sense of this whole expression often requires more effort than the difference between const T* and T* const. I feel that students need to know about const-overloading and how to do it without needlessly duplicating the code in the two functions, but this trick seems a bit excessive in an introductory C++ programming course. My question is this - is there a simpler way to implement const-overloaded functions in terms of one another? Or is there a simpler way of explaining this existing trick to students? Thanks so much!

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• Java: Making concurrent MySQL queries from multiple clients synchronised

  - by Misha Gale

  I work at a gaming cybercafe, and we've got a system here (smartlaunch) which keeps track of game licenses. I've written a program which interfaces with this system (actually, with it's backend MySQL database). The program is meant to be run on a client PC and (1) query the database to select an unused license from the pool available, then (2) mark this license as in use by the client PC. The problem is, I've got a concurrency bug. The program is meant to be launched simultaneously on multiple machines, and when this happens, some machines often try and acquire the same license. I think that this is because steps (1) and (2) are not synchronised, i.e. one program determines that license #5 is available and selects it, but before it can mark #5 as in use another copy of the program on another PC tries to grab that same license. I've tried to solve this problem by using transactions and table locking, but it doesn't seem to make any difference - Am I doing this right? Here follows the code in question: public LicenseKey Acquire() throws SmartLaunchException, SQLException { Connection conn = SmartLaunchDB.getConnection(); int PCID = SmartLaunchDB.getCurrentPCID(); conn.createStatement().execute("LOCK TABLE `licensekeys` WRITE"); String sql = "SELECT * FROM `licensekeys` WHERE `InUseByPC` = 0 AND LicenseSetupID = ? ORDER BY `ID` DESC LIMIT 1"; PreparedStatement statement = conn.prepareStatement(sql); statement.setInt(1, this.id); ResultSet results = statement.executeQuery(); if (results.next()) { int licenseID = results.getInt("ID"); sql = "UPDATE `licensekeys` SET `InUseByPC` = ? WHERE `ID` = ?"; statement = conn.prepareStatement(sql); statement.setInt(1, PCID); statement.setInt(2, licenseID); statement.executeUpdate(); statement.close(); conn.commit(); conn.createStatement().execute("UNLOCK TABLES"); return new LicenseKey(results.getInt("ID"), this, results.getString("LicenseKey"), results.getInt("LicenseKeyType")); } else { throw new SmartLaunchException("All licenses of type " + this.name + "are in use"); } }

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• Sync services not actually syncing

  - by Paul Mrozowski

  I'm attempting to sync a SQL Server CE 3.5 database with a SQL Server 2008 database using MS Sync Services. I am using VS 2008. I created a Local Database Cache, connected it with SQL Server 2008 and picked the tables I wanted to sync. I selected SQL Server Tracking. It modified the database for change tracking and created a local copy (SDF) of the data. I need two way syncing so I created a partial class for the sync agent and added code into the OnInitialized() to set the SyncDirection for the tables to Bidirectional. I've walked through with the debugger and this code runs. Then I created another partial class for cache server sync provider and added an event handler into the OnInitialized() to hook into the ApplyChangeFailed event. This code also works OK - my code runs when there is a conflict. Finally, I manually made some changes to the server data to test syncing. I use this code to fire off a sync: var agent = new FSEMobileCacheSyncAgent(); var syncStats = agent.Synchronize(); syncStats seems to show the count of the # of changes I made on the server and shows that they were applied. However, when I open the local SDF file none of the changes are there. I basically followed the instructions I found here: http://msdn.microsoft.com/en-us/library/cc761546%28SQL.105%29.aspx and here: http://keithelder.net/blog/archive/2007/09/23/Sync-Services-for-SQL-Server-Compact-Edition-3.5-in-Visual.aspx It seems like this should "just work" at this point, but the changes made on the server aren't in the local SDF file. I guess I'm missing something but I'm just not seeing it right now. I thought this might be because I appeared to be using version 1 of Sync Services so I removed the references to Microsoft.Synchronization.* assemblies, installed the Sync framework 2.0 and added the new version of the assemblies to the project. That hasn't made any difference. Ideas? Edit: I wanted to enable tracing to see if I could track this down but the only way to do that is through a WinForms app since it requires entries in the app.config file (my original project was a class library). I created a WinForms project and recreated everything and suddenly everything is working. So apparently this requires a WinForm project for some reason? This isn't really how I planned on using this - I had hoped to kick off syncing through another non-.NET application and provide the UI there so the experience was a bit more seemless to the end user. If I can't do that, that's OK, but I'd really like to know if/how to make this work as a class library project instead.

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• How can a C/C++ program put itself into background?

  - by Larry Gritz

  What's the best way for a running C or C++ program that's been launched from the command line to put itself into the background, equivalent to if the user had launched from the unix shell with '&' at the end of the command? (But the user didn't.) It's a GUI app and doesn't need any shell I/O, so there's no reason to tie up the shell after launch. But I want a shell command launch to be auto-backgrounded without the '&' (or on Windows). Ideally, I want a solution that would work on any of Linux, OS X, and Windows. (Or separate solutions that I can select with #ifdef.) It's ok to assume that this should be done right at the beginning of execution, as opposed to somewhere in the middle. One solution is to have the main program be a script that launches the real binary, carefully putting it into the background. But it seems unsatisfying to need these coupled shell/binary pairs. Another solution is to immediately launch another executed version (with 'system' or CreateProcess), with the same command line arguments, but putting the child in the background and then having the parent exit. But this seems clunky compared to the process putting itself into background. Edited after a few answers: Yes, a fork() (or system(), or CreateProcess on Windows) is one way to sort of do this, that I hinted at in my original question. But all of these solutions make a SECOND process that is backgrounded, and then terminate the original process. I was wondering if there was a way to put the EXISTING process into the background. One difference is that if the app was launched from a script that recorded its process id (perhaps for later killing or other purpose), the newly forked or created process will have a different id and so will not be controllable by any launching script, if you see what I'm getting at. Edit #2: fork() isn't a good solution for OS X, where the man page for 'fork' says that it's unsafe if certain frameworks or libraries are being used. I tried it, and my app complains loudly at runtime: "The process has forked and you cannot use this CoreFoundation functionality safely. You MUST exec()." I was intrigued by daemon(), but when I tried it on OS X, it gave the same error message, so I assume that it's just a fancy wrapper for fork() and has the same restrictions. Excuse the OS X centrism, it just happens to be the system in front of me at the moment. But I am indeed looking for a solution to all three platforms.

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