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

  - by Matma

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

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

  - by runrunraygun

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

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• Showing login view controller before main tab bar controller

  - by Padawan

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

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• Please Explain Drupal schema and drupal_write_record

  - by Aaron

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

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• AutoScaleMode problems with changed default font

  - by Doc Brown

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

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• BindException with INTERNET permission requested

  - by Mondain

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

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• Google maps API - info window height and panning

  - by Tim Fountain

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

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• Haskell: Left-biased/short-circuiting function

  - by user2967411

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

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• Asp.net MVC and MOSS 2010 integration

  - by Robert Koritnik

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

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

  - by marcgg

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

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

  - by TJ Thompson

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

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• Opera Mobile, offline web app development, and memory

  - by Jake Krohn

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

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• Problems related to showing MessageBox from non-GUI threads

  - by Hans Løken

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

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• When is it worth using a BindingSource?

  - by Justin

  I think I understand well enough what the BindingSource class does - i.e. provide a layer of indirection between a data source and a UI control. It implements the IBindingList interface and therefore also provides support for sorting. And I've used it frequently enough, without too many problems. But I'm wondering if I use it more often than I should. Perhaps an example would help. Let's say I have just a simple textbox on a form (using WinForms), and I'd like to bind that textbox to a simple property inside a class that returns a string. Is it worth using a BindingSource in this situation? Now let's say I have a grid on my form, and I'd like to bind it to a DataTable. Should I use a BindingSource now? In the latter case, I probably would not use a BindingSource, as a DataTable, from what I can gather, provides the same functionality that the BindingSource itself would. The DataTable will fire the the right events when a row is added, deleted, etc so that the grid will automatically update. But in the first case with the textbox being bound to a string, I would probably have the class that contains the string property implement INotifyPropertyChanged, so that it could fire the PropertyChanged event when the string changes. I would use a BindingSource so that it could listen to these PropertyChanged events so that it could update the textbox automatically when the string changes. How does this sound so far? I still feel like there's a gap in my understanding that's preventing me from seeing the whole picture. This has been a pretty vague question so far, so I'll try to ask some more specific questions - ideally the answers will reference the above examples or something similar... (1) Is it worth using a BindingSource in either of the above examples? (2) It seems that developers just "assume" that the DataTable class will do the right thing, in firing PropertyChanged events at the right time. How does one know if a data source is capable of doing this? Is there a particular interface that a data source should implement in order for developers to be able to assume this behaviour? (3) Does it matter what Control is being bound to, when considering whether or not to use a BindingSource? Or is it only the data source that should affect the decision? Perhaps the answer is (and this would seem logical enough): the Control needs to be intelligent enough to listen to the PropertyChanged events, otherwise a BindingSource is required. So how does one tell if the Control is capable of doing this? Again, is there a particular interface that developers can look for that the Control must implement? It is this confusion that has, in the past, led to me always using a BindingSource. But I'd like to understand better exactly when to use one, so that I do so only when necessary.

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

  - by mikep

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

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• Populating ComboBoxDataColumn items and values

  - by MarceloRamires

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

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• How can I take the first 100 characters of html content ( without stripping the TAGS! )

  - by Atomiton

  There are lots of questions on how to strip html tags, but not many on functions/methods to close them. Here's the situation. I have a 500 character Message summary ( which includes html tags ), but I only want the first 100 characters. Problem is if I truncate the message, it could be in the middle of an html tag... which messes up stuff. Assuming the html is something like this: <div class="bd">"Lorem ipsum dolor sit amet, consectetur adipisicing elit, sed do eiusmod tempor incididunt ut labore et dolore magna aliqua. <br/> <br/>Some Dates: April 30 - May 2, 2010 <br/> <p>Ut enim ad minim veniam, quis nostrud exercitation ullamco laboris nisi ut aliquip ex ea commodo consequat. <em>Duis aute irure dolor in reprehenderit</em> in voluptate velit esse cillum dolore eu fugiat nulla pariatur. Excepteur sint occaecat cupidatat non proident, sunt in culpa qui officia deserunt mollit anim id est laborum. <br/> </p> For more information about Lorem Ipsum doemdloe, visit: <br/> <a href="http://www.somesite.com" title="Some Conference">Some text link</a><br/> </div> How would I take the first ~100 characters or so? ( Although, ideally that would be the first approximately 100 characters of "CONTENT" ( in between the html tags ) I'm assuming the best way to do this would be a recursive algorithm that keeps track of the html tags and appends any tags that would be truncated, but that may not be the best approach. My first thoughts are using recursion to count nested tags, and when we reach 100 characters, look for the next "<" and then use recursion to write the closing html tags needed from there. The reason for doing this is to make a short summary of existing articles without requiring the user to go back and provide summaries for all the articles. I want to keep the html formatting, if possible. NOTE: Please ignore that the html isn't totally semantic. This is what I have to deal with from my WYSIWYG. EDIT: I added a potential solution ( that seems to work ) I figure others will run into this problem as well. I'm not sure it's the best... and it's probably not totally robust ( in fact, I know it isn't ), but I'd appreciate any feedback

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• Mixing application modules between Silverlight and ASP.NET

  - by jkohlhepp

  Background: I work in a suite of ASP.NET applications that have several different "modules". The applications all share a main menu, so they all link to one-another. The modules are the high-level areas of the application. So, for example, it might be Payments, Orders, Customers, Products, etc. And Payments and Orders are in one app and Products and Customers are in another. Some of these menu links are "deep links", for example it might be a link to a particular page within the Customers module, such as Create New Customer. The issue: We are about to start a project that will add several more modules to this suite, probably as a new .NET application. I'm thinking about doing these new modules in Silverlight (for various reasons that are not material to the question). If I were to do that, I need to make the menu look the same as the menu in ASP.NET, as the users still need to feel like they are inside one "application". My questions: How should I organize the Silverlight project(s) so that I can "deep link" from ASP.NET pages into particular modules in the Silverlight app? What is even the best idea for creating these different Silverlight "modules"? If I had something that would've been a page in ASP.NET (for example - Create Customer), should each one of those be a separate Silverlight app? Or should it be a separate User Control? Or something else? Should I reuse our shared ASP.NET menu, and deep link to different Silverlight "modules" even within the new application? Or should I reimplement the menu in Silverlight for navigation within the app? Are there menu controls for Silverlight that look similar to ASP.NET menus (with flyout submenus in this case)? Could I maybe even share a SiteMap XML file between them? Edit: After looking around a bit more, it seems like PRISM might be the answer for some of my issues. It would allow me to modularize the different chunks of Silverlight that I have. And it would allow me to define a "master page" in Silverlight where I could host the menu. Do I have this right?

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• Calculate year for end date: PostgreSQL

  - by Dave Jarvis

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

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• How to Practice Unix Programming in C?

  - by danben

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

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• Cascade Saves with Fluent NHibernate AutoMapping - Old Anwser Still Valid?

  - by Glenn

  I want to do exactly what this question asks: http://stackoverflow.com/questions/586888/cascade-saves-with-fluent-nhibernate-automapping Using Fluent Nhibernate Mappings to turn on "cascade" globally once for all classes and relation types using one call rather than setting it for each mapping individually. The answer to the earlier question looks great, but I'm afraid that the Fluent Nhibernate API altered its .WithConvention syntax last year and broke the answer... either that or I'm missing something. I keep getting a bunch of name space not found errors relating to the IOneToOnePart, IManyToOnePart and all their variations: "The type or namespace name 'IOneToOnePart' could not be found (are you missing a using directive or an assembly reference?)" I've tried the official example dll's, the RTM dll's and the latest build and none of them seem to make VS 2008 see the required namespace. The second problem is that I want to use the class with my AutoPersistenceModel but I'm not sure where to this line: .ConventionDiscovery.AddFromAssemblyOf() in my factory creation method. private static ISessionFactory CreateSessionFactory() { return Fluently.Configure() .Database(SQLiteConfiguration.Standard.UsingFile(DbFile)) .Mappings(m => m.AutoMappings .Add(AutoMap.AssemblyOf<Shelf>(type => type.Namespace.EndsWith("Entities")) .Override<Shelf>(map => { map.HasManyToMany(x => x.Products).Cascade.All(); }) ) )//emd mappings .ExposeConfiguration(BuildSchema) .BuildSessionFactory();//finalizes the whole thing to send back. } Below is the class and using statements I'm trying using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.IO; using FluentNHibernate.Conventions; using FluentNHibernate.Cfg; using FluentNHibernate.Cfg.Db; using NHibernate; using NHibernate.Cfg; using NHibernate.Tool.hbm2ddl; using FluentNHibernate.Mapping; namespace TestCode { public class CascadeAll : IHasOneConvention, IHasManyConvention, IReferenceConvention { public bool Accept(IOneToOnePart target) { return true; } public void Apply(IOneToOnePart target) { target.Cascade.All(); } public bool Accept(IOneToManyPart target) { return true; } public void Apply(IOneToManyPart target) { target.Cascade.All(); } public bool Accept(IManyToOnePart target) { return true; } public void Apply(IManyToOnePart target) { target.Cascade.All(); } } }

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• PriorityQueue update problems

  - by Bharat

  After going through a bunch of questions here on SO, I still have no idea what exactly is going wrong with my code and would appreciate some help. I'm trying to implement a priority queue based on f-costs for an A* algorithm, and while the algorithm works fine for short pathfinding distances, it seems to go wrong when there's an obstacle or when the distance between start and goal points is greater than about 30 squares (although sometimes it screws up for less too). while(!m_qOpenList.isEmpty()) { m_xCurrent=m_qOpenList.poll(); m_xCurrent.setBackground(Color.red); m_qClosedList.add(m_xCurrent); if(m_xCurrent.getStatus()==2) { System.out.println("Target Reached"); solved=true; break; } iX=m_xCurrent.getXCo(); iY=m_xCurrent.getYCo(); for(i=iX-1;i<=iX+1;i++) for(j=iY-1;j<=iY+1;j++) { if(i<0||j<0||i>m_iMazeX||j>m_iMazeX||(i==iX&&j==iY) || m_xNode[i][j].getStatus()==4|| m_qClosedList.contains(m_xNode[i][j])) continue; m_xNode[i][j].score(m_xCurrent,m_xGoal); m_qOpenList.add(m_xNode[i][j]); } } It's quite rudimentary as I'm just trying to get it to work for now. m_qOpenList is the PriorityQueue. The problem is that when I debug the program, at some point (near an obstacle), a Node with a fcost of say 84 has higher priority than a node with an fcost of 70. I am not attempting to modify the values once they're on the priority queue. You'll notice that I add at the end of the while loop (I read somewhere that the priorityqueue reorders itself when stuff is added to it), and poll right after that at the beginning. Status of 2 means the Node is the goal, and a status of 4 means that it is unwalkable. public int compareTo(Node o) { if(m_iF<o.m_iF) return -1; if(m_iF>o.m_iF) return 1; return 0; } And that's the compareTo function. Can you see a problem? =(

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• Git for Websites / post-receive / Separation of Test and Production Sites

  - by Walt W

  Hi all, I'm using Git to manage my website's source code and deployment, and currently have the test and live sites running on the same box. Following this resource http://toroid.org/ams/git-website-howto originally, I came up with the following post-receive hook script to differentiate between pushes to my live site and pushes to my test site: while read ref do #echo "Ref updated:" #echo $ref -- would print something like example at top of file result=`echo $ref | gawk -F' ' '{ print $3 }'` if [ $result != "" ]; then echo "Branch found: " echo $result case $result in refs/heads/master ) git --work-tree=c:/temp/BLAH checkout -f master echo "Updated master" ;; refs/heads/testbranch ) git --work-tree=c:/temp/BLAH2 checkout -f testbranch echo "Updated testbranch" ;; * ) echo "No update known for $result" ;; esac fi done echo "Post-receive updates complete" However, I have doubts that this is actually safe :) I'm by no means a Git expert, but I am guessing that Git probably keeps track of the current checked-out branch head, and this approach probably has the potential to confuse it to no end. So a few questions: IS this safe? Would a better approach be to have my base repository be the test site repository (with corresponding working directory), and then have that repository push changes to a new live site repository, which has a corresponding working directory to the live site base? This would also allow me to move the production to a different server and keep the deployment chain intact. Is there something I'm missing? Is there a different, clean way to differentiate between test and production deployments when using Git for managing websites? As an additional note in light of Vi's answer, is there a good way to do this that would handle deletions without mucking with the file system much? Thank you, -Walt PS - The script I came up with for the multiple repos (and am using unless I hear better) is as follows: sitename=`basename \`pwd\`` while read ref do #echo "Ref updated:" #echo $ref -- would print something like example at top of file result=`echo $ref | gawk -F' ' '{ print $3 }'` if [ $result != "" ]; then echo "Branch found: " echo $result case $result in refs/heads/master ) git checkout -q -f master if [ $? -eq 0 ]; then echo "Test Site checked out properly" else echo "Failed to checkout test site!" fi ;; refs/heads/live-site ) git push -q ../Live/$sitename live-site:master if [ $? -eq 0 ]; then echo "Live Site received updates properly" else echo "Failed to push updates to Live Site" fi ;; * ) echo "No update known for $result" ;; esac fi done echo "Post-receive updates complete" And then the repo in ../Live/$sitename (these are "bare" repos with working trees added after init) has the basic post-receive: git checkout -f if [ $? -eq 0 ]; then echo "Live site `basename \`pwd\`` checked out successfully" else echo "Live site failed to checkout" fi

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• SilverLight 3 Beginner question: Scroll with mousewheel and zoom image with panning

  - by JP Hellemons

  Hello, I would like to make a small silverlight app which displays one fairly large image which can be zoomed in by scrolling the mouse and then panned with the mouse. it's similar to the function in google maps and i do not want to use deepzoom. here is what i have at the moment. please keep in mind that this is my first silverlight app: this app is just for me to see it's a good way to build in a website. so it's a demo app and therefor has bad variable names. the initial image is 1800px width. private void sc_MouseWheel(object sender, MouseWheelEventArgs e) { var st = (ScaleTransform)plaatje.RenderTransform; double zoom = e.Delta > 0 ? .1 : -.1; st.ScaleX += zoom; st.ScaleY += zoom; } this works, but could use some smoothing and it's positioned top left and not centered. the panning is like this: found it @ http://stackoverflow.com/questions/741956/wpf-pan-zoom-image and converted it to this below to work in silverlight Point start; Point origin; bool captured = false; private void plaatje_MouseLeftButtonDown(object sender, MouseButtonEventArgs e) { plaatje.CaptureMouse(); captured = true; var tt = (TranslateTransform)((TransformGroup)plaatje.RenderTransform) .Children.First(tr => tr is TranslateTransform); start = e.GetPosition(canvasje); origin = new Point(tt.X, tt.Y); } private void plaatje_MouseLeftButtonUp(object sender, MouseButtonEventArgs e) { plaatje.ReleaseMouseCapture(); captured = false; } private void plaatje_MouseMove(object sender, MouseEventArgs e) { if (!captured) return; var tt = (TranslateTransform)((TransformGroup)plaatje.RenderTransform).Children.First(tr => tr is TranslateTransform); double xVerschuiving = start.X - e.GetPosition(canvasje).X; double yVerschuiving = start.Y - e.GetPosition(canvasje).Y; tt.X = origin.X - xVerschuiving; tt.Y = origin.Y - yVerschuiving; } so the scaling isn't smooth and the panning isn't working, because when i click it, the image disappears. thanks in advanced!

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• Succinct introduction to C++/CLI for C#/Haskell/F#/JS/C++/... programmer

  - by Henrik

  Hello everybody, I'm trying to write integrations with the operating system and with things like active directory and Ocropus. I know a bunch of programming languages, including those listed in the title. I'm trying to learn exactly how C++/CLI works, but can't find succinct, exact and accurate descriptions online from the searching that I have done. So I ask here. Could you tell me the pitfalls and features of C++/CLI? Assume I know all of C# and start from there. I'm not an expert in C++, so some of my questions' answers might be "just like C++", but could say that I am at C#. I would like to know things like: Converting C++ pointers to CLI pointers, Any differences in passing by value/doubly indirect pointers/CLI pointers from C#/C++ and what is 'recommended'. How do gcnew, __gc, __nogc work with Polymorphism Structs Inner classes Interfaces The "fixed" keyword; does that exist? Compiling DLLs loaded into the kernel with C++/CLI possible? Loaded as device drivers? Invoked by the kernel? What does this mean anyway (i.e. to load something into the kernel exactly; how do I know if it is?)? L"my string" versus "my string"? wchar_t? How many types of chars are there? Are we safe in treating chars as uint32s or what should one treat them as to guarantee language indifference in code? Finalizers (~ClassName() {}) are discouraged in C# because there are no garantuees they will run deterministically, but since in C++ I have to use "delete" or use copy-c'tors as to stack allocate memory, what are the recommendations between C#/C++ interactions? What are the pitfalls when using reflection in C++/CLI? How well does C++/CLI work with the IDisposable pattern and with SafeHandle, SafeHandleZeroOrMinusOneIsInvalid? I've read briefly about asynchronous exceptions when doing DMA-operations, what are these? Are there limitations you impose upon yourself when using C++ with CLI integration rather than just doing plain C++? Attributes in C++ similar to Attributes in C#? Can I use the full meta-programming patterns available in C++ through templates now and still have it compile like ordinary C++? Have you tried writing C++/CLI with boost? What are the optimal ways of interfacing the boost library with C++/CLI; can you give me an example of passing a lambda expression to an iterator/foldr function? What is the preferred way of exception handling? Can C++/CLI catch managed exceptions now? How well does dynamic IL generation work with C++/CLI? Does it run on Mono? Any other things I ought to know about?

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