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  • Game doesn't Quit properly

    - by W.K.S
    I have an app that so far consists of two Activities: The Main Menu Activity. The Game Activity The Main Menu Activity contains a button that starts the Game Activity with the following code: public void onClick(View clickedButton) { switch(clickedButton.getId()) { case R.id.buttonPlay: Intent i = new Intent("apple.banana.BouncingBallActivity"); startActivity(i); break; } When the user is done with the Game Activity, he presses the back button. This calls the onPause() method first, which pauses the animation thread of the game. It then calls the onStop() which calls finish() on the activity altogether. The user is returned to the Main Menu activity. The code is outlined below: public class BouncingBallActivity extends Activity{ private BouncingBallView bouncingBallView; @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); bouncingBallView = new BouncingBallView(this); bouncingBallView.resume(); setContentView(bouncingBallView); } @Override protected void onPause() { super.onPause(); bouncingBallView.pause(); } @Override protected void onResume() { super.onResume(); bouncingBallView.resume(); } @Override protected void onStop() { super.onStop(); this.finish(); } } The problem is that this only works if I launch the application from Eclipse. When I click on the app icon, the game starts from the Game Activity. The main menu activity does not appear. I am not clear about why this happens. It could be something to do with the manifest. I've pasted the relevant portions below: <application android:icon="@drawable/ic_launcher" android:label="@string/app_name" > <activity android:name=".BouncingBallActivity" android:label="@string/app_name" android:screenOrientation="landscape" > <intent-filter> <action android:name="apple.banana.BouncingBallActivity" /> <category android:name="android.intent.category.DEFAULT" /> </intent-filter> </activity> <activity android:name=".MainMenu" android:label="@string/app_name" android:screenOrientation="portrait" > <intent-filter> <action android:name="android.intent.action.MAIN" /> <category android:name="android.intent.category.LAUNCHER" /> </intent-filter> </activity> </application> I'd really appreciate any help with this. Thanks.

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  • Passing custom info to mongrel_rails start

    - by whaka
    One thing I really don't understand is how I can pass custom start-up options to a mongrel instance. I see that a common approach is the use environment variables, but in my environment this is not going to work because my rails application serves many different clients. Much code is shared between clients, but there are also many differences which I implement by subclassing controllers and views to overload or extend existing features or introduce new ones. To make this all work, I simply add the paths to client specific modules the module load path ($:). In order to start the application for a particular client, I could now use an environment variable like say, TARGET=AMAZONE. Unfortunately, on some systems I'm running multiple mongrel clusters, each cluster serving a different client. Some of these systems run under Windows and to start mongrel I installed mongrel_services. Clearly, this makes my environment variable unsuitable. Passing this extra bit of data to the application is proving to be a real challenge. For a start, mongrel_rails service_install will reject any [custom] command line parameters that aren't documented. I'm not too concerned as installing the services using the install program is trivial. However, even if I manage to install mongrel_services such that when run it passes the custom command line option --target to mongrel_rails start, I get an error because mongrel_rails doesn't recognize the switch. So here were the things I looked at: Pass an extra parameter: mongrel_rails start --target XYZ ... use a config file and add target:XYZ, then do: mongrel_rails start -C x:\myapp\myconfig.yml modify the file: Ruby\lib\ruby\gems\1.8\gems\mongrel-1.1.5-x86-mswin32-60\lib\mongrel\command.rb Perhaps I can use the --script option, but all docs that I found on it were for Unix 1 and 2 simply don't work. I played with 4 but never managed it to do anything. So I had no choice but to go with 3. While it is relatively simple, I hate changing ruby library code. Particularly disappointing is that 2 doesn't work. I mean what is so unreasonable about adding other [custom] options in the config file? Actually I think this is a fundamental piece that is missing in rails. Somehow, the application should be able to register and access command line arguments it expects. If anybody has a good idea how to do this more elegantly using the current infrastructure, I have a chocolate fish to give away!!!

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  • Issue when rotating a UIScrollView

    - by leachianus.gecko
    I am having issues trying to get the pageControl sample code to work with rotation. I managed to get it to rotate but it does not visually load correctly until I start to scroll (then it works fine). Any Idea on how I can fix this problem? Here is a link to the project if you want to see it in action. This code is based off the PageControl example apple has provided. here is the code: #import "ScrollingViewController.h" #import "MyViewController.h" @interface ScrollingViewController (PrivateMethods) - (void)loadScrollViewWithPage:(int)page; @end @implementation ScrollingViewController @synthesize scrollView; @synthesize viewControllers; - (void)viewDidLoad { amount = 5; [super viewDidLoad]; [self setupPage]; } - (void)didReceiveMemoryWarning { [super didReceiveMemoryWarning]; } - (void)viewDidUnload { [scrollView release]; } - (void)dealloc { [super dealloc]; } - (void)setupPage { NSMutableArray *controllers = [[NSMutableArray alloc] init]; for (unsigned i = 0; i < amount; i++) { [controllers addObject:[NSNull null]]; } self.viewControllers = controllers; [controllers release]; // a page is the width of the scroll view scrollView.pagingEnabled = YES; scrollView.contentSize = CGSizeMake(scrollView.frame.size.width * amount, 200); scrollView.showsHorizontalScrollIndicator = NO; scrollView.showsVerticalScrollIndicator = NO; scrollView.scrollsToTop = NO; scrollView.delegate = self; [self loadScrollViewWithPage:0]; [self loadScrollViewWithPage:1]; } #pragma mark - #pragma mark UIScrollViewDelegate stuff - (void)scrollViewDidScroll:(UIScrollView *)_scrollView { if (pageControlIsChangingPage) { return; } /* * We switch page at 50% across */ CGFloat pageWidth = _scrollView.frame.size.width; int dog = floor((_scrollView.contentOffset.x - pageWidth / 2) / pageWidth) + 1; // pageControl.currentPage = page; [self loadScrollViewWithPage:dog - 1]; [self loadScrollViewWithPage:dog]; [self loadScrollViewWithPage:dog + 1]; } - (void)loadScrollViewWithPage:(int)page { if (page < 0) return; if (page >= amount) return; MyViewController *controller = [viewControllers objectAtIndex:page]; if ((NSNull *)controller == [NSNull null]) { controller = [[MyViewController alloc] initWithPageNumber:page]; [viewControllers replaceObjectAtIndex:page withObject:controller]; [controller release]; } if (nil == controller.view.superview) { CGRect frame = scrollView.frame; frame.origin.x = frame.size.width * page; frame.origin.y = 0; controller.view.frame = frame; [scrollView addSubview:controller.view]; } } - (void)didRotateFromInterfaceOrientation:(UIInterfaceOrientation)fromInterfaceOrientation { [self setupPage]; } - (BOOL)shouldAutorotateToInterfaceOrientation:(UIInterfaceOrientation)interfaceOrientation { // Return YES for supported orientations return YES; } @end

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  • avoiding enums as interface identifiers c++ OOP

    - by AlasdairC
    Hi I'm working on a plugin framework using dynamic loaded shared libraries which is based on Eclipse's (and probally other's) extension-point model. All plugins share similar properties (name, id, version etc) and each plugin could in theory satisfy any extension-point. The actual plugin (ie Dll) handling is managed by another library, all I am doing really is managing collections of interfaces for the application. I started by using an enum PluginType to distinguish the different interfaces, but I have quickly realised that using template functions made the code far cleaner and would leave the grunt work up to the compiler, rather than forcing me to use lots of switch {...} statements. The only issue is where I need to specify like functionality for class members - most obvious example is the default plugin which provides a particular interface. A Settings class handles all settings, including the default plugin for an interface. ie Skin newSkin = settings.GetDefault<ISkin>(); How do I store the default ISkin in a container without resorting to some other means of identifying the interface? As I mentioned above, I currently use a std::map<PluginType, IPlugin> Settings::defaults member to achieve this (where IPlugin is an abstract base class which all plugins derive from. I can then dynamic_cast to the desired interface when required, but this really smells of bad design to me and introduces more harm than good I think. would welcome any tips edit: here's an example of the current use of default plugins typedef boost::shared_ptr<ISkin> Skin; typedef boost::shared_ptr<IPlugin> Plugin; enum PluginType { skin, ..., ... } class Settings { public: void SetDefault(const PluginType type, boost::shared_ptr<IPlugin> plugin) { m_default[type] = plugin; } boost::shared_ptr<IPlugin> GetDefault(const PluginType type) { return m_default[type]; } private: std::map<PluginType, boost::shared_ptr<IPlugin> m_default; }; SkinManager::Initialize() { Plugin thedefault = g_settings.GetDefault(skinplugin); Skin defaultskin = boost::dynamic_pointer_cast<ISkin>(theskin); defaultskin->Initialize(); } I would much rather call the getdefault as the following, with automatic casting to the derived class. However I need to specialize for every class type. template<> Skin Settings::GetDefault<ISkin>() { return boost::dynamic_pointer_cast<ISkin>(m_default(skin)); }

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  • jqgrid modify data returned from ajax call before display in table

    - by user954211
    I have to display some data that I receive from the server as json object like this {"rowndx":"0","rows":"25","rowstotal":"100","rowsdata":[ ["00","DEVICE001","T0_IHOME","1","***","1","10"], ["01","DEVICE002","NO_DEVICE","1","***","1","10"], ["02","DEVICE003","NO_DEVICE","0","***","1","10"], ..... Before displaying the received data in a table I would like to make changes where necessary adding units to the numbers or replacing the numbers with words (eg 0 -OFF 1- ON) To do this I have associated at the ajax option "success" my encoding function. In this case, however, remains always visible the message "Loading ..." and no other action is permitted. I moved my re-encoding procedure to the "complete" ajax option and this time it seems to work. But I did not understand what was my mistake and I do not know if my procedure can work. This is my table ajax configuration url : "devtbl.json", mtype : "POST", datatype : "json", postData : ...... ajaxGridOptions: { type : 'post', contentType: 'application/json', async : false, complete : DEVparse_serverdata, error : function() { alert('Something bad happened. Stopping');}, }, jsonReader : { root : "tablerows", page : "currentpage", total : "totalpages", records : "totalrecords", cell : "", id : "0", userdata : "userdata", repeatitems : true }, and my coding function function DEVparse_serverdata(js , textStatus) { var jsontablereply = {} ; var rowsxpage_int = parseInt(UB.rowsxpage.DEVtable) ; var jsonreply = jQuery.parseJSON(js.responseText) ; jsontablereply.currentpage = "" + (1 + (parseInt(jsonreply.rowndx) / rowsxpage_int)); jsontablereply.totalpages = "" + parseInt((parseInt(jsonreply.rowstotal) + (rowsxpage_int-1)) / rowsxpage_int) ; jsontablereply.totalrecords = jsonreply.rowstotal; jsontablereply.tablerows = [] ; $.each(jsonreply.rowsdata, function(ndx, row) { var rowarray = [] ; rowarray[0] = row[0] ; rowarray[1] = row[1] ; rowarray[2] = row[2] ; rowarray[3] = row[3] ; rowarray[4] = row[4] ; switch (row[2]) { case "NO_DEVICE": rowarray[5] = "***" ; break ; case "T0_IHOME": rowarray[5] = "T=" + row[5] + "°C" ; break ; } jsontablereply.tablerows[ndx] = rowarray ; }) ; // each jQuery("#DEVtbl")[0].addJSONData(jsontablereply); } (I am a beginner with Jquery my coding style is poor)

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  • How to make MySQL utilize available system resources, or find "the real problem"?

    - by anonymous coward
    This is a MySQL 5.0.26 server, running on SuSE Enterprise 10. This may be a Serverfault question. The web user interface that uses these particular queries (below) is showing sometimes 30+, even up to 120+ seconds at the worst, to generate the pages involved. On development, when the queries are run alone, they take up to 20 seconds on the first run (with no query cache enabled) but anywhere from 2 to 7 seconds after that - I assume because the tables and indexes involved have been placed into ram. From what I can tell, the longest load times are caused by Read/Update Locking. These are MyISAM tables. So it looks like a long update comes in, followed by a couple 7 second queries, and they're just adding up. And I'm fine with that explanation. What I'm not fine with is that MySQL doesn't appear to be utilizing the hardware it's on, and while the bottleneck seems to be the database, I can't understand why. I would say "throw more hardware at it", but we did and it doesn't appear to have changed the situation. Viewing a 'top' during the slowest times never shows much cpu or memory utilization by mysqld, as if the server is having no trouble at all - but then, why are the queries taking so long? How can I make MySQL use the crap out of this hardware, or find out what I'm doing wrong? Extra Details: On the "Memory Health" tab in the MySQL Administrator (for Windows), the Key Buffer is less than 1/8th used - so all the indexes should be in RAM. I can provide a screen shot of any graphs that might help. So desperate to fix this issue. Suffice it to say, there is legacy code "generating" these queries, and they're pretty much stuck the way they are. I have tried every combination of Indexes on the tables involved, but any suggestions are welcome. Here's the current Create Table statement from development (the 'experimental' key I have added, seems to help a little, for the example query only): CREATE TABLE `registration_task` ( `id` varchar(36) NOT NULL default '', `date_entered` datetime NOT NULL default '0000-00-00 00:00:00', `date_modified` datetime NOT NULL default '0000-00-00 00:00:00', `assigned_user_id` varchar(36) default NULL, `modified_user_id` varchar(36) default NULL, `created_by` varchar(36) default NULL, `name` varchar(80) NOT NULL default '', `status` varchar(255) default NULL, `date_due` date default NULL, `time_due` time default NULL, `date_start` date default NULL, `time_start` time default NULL, `parent_id` varchar(36) NOT NULL default '', `priority` varchar(255) NOT NULL default '9', `description` text, `order_number` int(11) default '1', `task_number` int(11) default NULL, `depends_on_id` varchar(36) default NULL, `milestone_flag` varchar(255) default NULL, `estimated_effort` int(11) default NULL, `actual_effort` int(11) default NULL, `utilization` int(11) default '100', `percent_complete` int(11) default '0', `deleted` tinyint(1) NOT NULL default '0', `wf_task_id` varchar(36) default '0', `reg_field` varchar(8) default '', `date_offset` int(11) default '0', `date_source` varchar(10) default '', `date_completed` date default '0000-00-00', `completed_id` varchar(36) default NULL, `original_name` varchar(80) default NULL, PRIMARY KEY (`id`), KEY `idx_reg_task_p` (`deleted`,`parent_id`), KEY `By_Assignee` (`assigned_user_id`,`deleted`), KEY `status_assignee` (`status`,`deleted`), KEY `experimental` (`deleted`,`status`,`assigned_user_id`,`parent_id`,`date_due`) ) ENGINE=MyISAM DEFAULT CHARSET=latin1 And one of the ridiculous queries in question: SELECT users.user_name assigned_user_name, registration.FIELD001 parent_name, registration_task.status status, registration_task.date_modified date_modified, registration_task.date_due date_due, registration.FIELD240 assigned_wf, if(LENGTH(registration_task.description)>0,1,0) has_description, registration_task.* FROM registration_task LEFT JOIN users ON registration_task.assigned_user_id=users.id LEFT JOIN registration ON registration_task.parent_id=registration.id where (registration_task.status != 'Completed' AND registration.FIELD001 LIKE '%' AND registration_task.name LIKE '%' AND registration.FIELD060 LIKE 'GN001472%') AND registration_task.deleted=0 ORDER BY date_due asc LIMIT 0,20; my.cnf - '[mysqld]' section. [mysqld] port = 3306 socket = /var/lib/mysql/mysql.sock skip-locking key_buffer = 384M max_allowed_packet = 100M table_cache = 2048 sort_buffer_size = 2M net_buffer_length = 100M read_buffer_size = 2M read_rnd_buffer_size = 160M myisam_sort_buffer_size = 128M query_cache_size = 16M query_cache_limit = 1M EXPLAIN above query, without additional index: +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ | id | select_type | table | type | possible_keys | key | key_len | ref | rows | Extra | +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ | 1 | SIMPLE | registration_task | ref | idx_reg_task_p,status_assignee | idx_reg_task_p | 1 | const | 1067354 | Using where; Using filesort | | 1 | SIMPLE | registration | eq_ref | PRIMARY,gbl | PRIMARY | 8 | sugarcrm401.registration_task.parent_id | 1 | Using where | | 1 | SIMPLE | users | ref | PRIMARY | PRIMARY | 38 | sugarcrm401.registration_task.assigned_user_id | 1 | | +----+-------------+-------------------+--------+--------------------------------+----------------+---------+------------------------------------------------+---------+-----------------------------+ EXPLAIN above query, with 'experimental' index: +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+ | id | select_type | table | type | possible_keys | key | key_len | ref | rows | Extra | +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+ | 1 | SIMPLE | registration_task | range | idx_reg_task_p,status_assignee,NewIndex1,tcg_experimental | tcg_experimental | 259 | NULL | 103345 | Using where; Using filesort | | 1 | SIMPLE | registration | eq_ref | PRIMARY,gbl | PRIMARY | 8 | sugarcrm401.registration_task.parent_id | 1 | Using where | | 1 | SIMPLE | users | ref | PRIMARY | PRIMARY | 38 | sugarcrm401.registration_task.assigned_user_id | 1 | | +----+-------------+-------------------+--------+-----------------------------------------------------------+------------------+---------+------------------------------------------------+--------+-----------------------------+

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  • Displaying individual elements of an object in an Arraylist through a for loop?

    - by user1180888
    I'm trying to Display individual elements of an Object I have created. It is a simple Java program that allows users to add and keep track of Player Details. I'm just stumped when it comes to displaying the details after they have been added already. here is what my code looks like I can create the object and input it into the arraylist no problem using the case 2, but when I try to print it out I want to do something like System.out.println("Player Name" + myPlayersArrayList.PlayerName + "Player Position" + myPlayerArrayList.PlayerPosition + "Player Age" + "Player Age"); I know that is not correct, but I dont really know what to do, if anyone can be of any help it would be greatly appreciated. Thanks System.out.println("Welcome to the Football Player database"); System.out.print(System.getProperty("line.separator")); UserInput myFirstUserInput = new UserInput(); int selection; ArrayList<Player> myPlayersArrayList = new ArrayList<Player>(); while (true) { System.out.println("1. View The Players"); System.out.println("2. Add A Player"); System.out.println("3. Edit A Player"); System.out.println("4. Delete A Player"); System.out.println("5. Exit ") ; System.out.print(System.getProperty("line.separator")); selection = myFirstUserInput.getInt("Please select an option"); System.out.print(System.getProperty("line.separator")); switch(selection){ case 1: if (myPlayersArrayList.isEmpty()) { System.out.println("No Players Have Been Entered Yet"); System.out.print(System.getProperty("line.separator")); break;} else {for(int i = 0; i < myPlayersArrayList.size(); i++){ System.out.println(myPlayersArrayList); } break; case 2: { String playerName,playerPos; int playerAge; playerName = (myFirstUserInput.getString("Enter Player name")); playerPos = (myFirstUserInput.getString("Enter Player Position")); playerAge = (myFirstUserInput.getInt("Enter Player Age")); myPlayersArrayList.add(new Player(playerName, playerPos, playerAge)); ; break; }

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  • jQuery 1.5.1 vs 1.4.4 weirdness

    - by zobgib
    I have been getting some weird errors when I upgrade jQuery from 1.4.4 to 1.5.1. Maybe you guys can explain what I need to change or why the new version is not working. In 1.4.4 I have the following code <div class="navlink home" data-link="home"> <span class="top">Home</span> </div> <div id="index-03"> </div> <div class="navlink resume" data-link="resume"> <span class="top">Resume</span> </div> <div id="index-05"> </div> <div id="index-06"> </div> <div class="navlink portfolio" data-link="portfolio"> <span class="bottom">Portfolio</span> </div> JS: $(".navlink").hover( function () { $(this).delay(100).animate({backgroundPosition: "-100% 0"}, 400); $(this).find("span").css("textDecoration","underline"); }, function () { $(this).queue("fx", []); $(this).animate({backgroundPosition: "0% 0%"}, 400); $(this).find("span").css("textDecoration","none"); } ); Which works just fine. but when I switch jQuery versions by changing this line in my header from <script type="text/javascript" src="https://ajax.googleapis.com/ajax/libs/jquery/1.4.4/jquery.min.js"></script> to <script type="text/javascript" src="https://ajax.googleapis.com/ajax/libs/jquery/1.5.1/jquery.min.js"></script> The above code quits animating and the background image just disappears. Here is a jsFiddle that shows what's happening just change the jQuery version on the side between 1.4.4 and 1.5.1 http://jsfiddle.net/fUXZ4/ -- 1.4.4 http://jsfiddle.net/3APCd/ -- 1.5.1 Here is a video of exactly what is happening to me: http://img.zobgib.com/2011-03-07_1905.swf

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  • Large memory chunk not garbage collected

    - by Niels
    In a hunt for a memory-leak in my app I chased down a behaviour I can't understand. I allocate a large memory block, but it doesn't get garbage-collected resulting in a OOM, unless I explicit null the reference in onDestroy. In this example I have two almost identical activities that switch between each others. Both have a single button. On pressing the button MainActivity starts OOMActivity and OOMActivity returns by calling finish(). After pressing the buttons a few times, Android throws a OOMException. If i add the the onDestroy to OOMActivity and explicit null the reference to the memory chunk, I can see in the log that the memory is correctly freed. Why doesn't the memory get freed automatically without the nulling? MainActivity: package com.example.oom; import android.app.Activity; import android.content.Intent; import android.os.Bundle; import android.view.View; import android.view.View.OnClickListener; import android.widget.Button; public class MainActivity extends Activity implements OnClickListener { private int buttonId; @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); System.gc(); Button OOMButton = new Button(this); OOMButton.setText("OOM"); buttonId = OOMButton.getId(); setContentView(OOMButton); OOMButton.setOnClickListener(this); } @Override public void onClick(View v) { if (v.getId() == buttonId) { Intent leakIntent = new Intent(this, OOMActivity.class); startActivity(leakIntent); } } } OOMActivity: public class OOMActivity extends Activity implements OnClickListener { private static final int WASTE_SIZE = 20000000; private byte[] waste; private int buttonId; protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); Button BackButton = new Button(this); BackButton.setText("Back"); buttonId = BackButton.getId(); setContentView(BackButton); BackButton.setOnClickListener(this); waste = new byte[WASTE_SIZE]; } public void onClick(View view) { if (view.getId() == buttonId) { finish(); } } }

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  • How can i use the Orientation correct for images

    - by user3578109
    I´m learning android/java by myself @the moment and i have a problem with a part of my app i´m learning on. I made the code with help of the www and my problem is that if i open an image from the gallery it´s send to the edit activity but in the activity pictures what are made in portrait mode are displayed always wrong (90° to the right side).... The codes are Matrix private Bitmap rotateBitmapToOrientation(Bitmap b, int orientation){ Matrix matrix = new Matrix(); matrix.postRotate(orientation); Canvas offscreenCanvas = new Canvas(); offscreenCanvas.drawBitmap(b, matrix, null); return b; } and the other one @Override protected void onActivityResult(int requestCode, int resultCode, Intent data) { super.onActivityResult(requestCode, resultCode, data); switch (requestCode) { case PICK_IMAGE_FROM_GALLERY: { if (resultCode == RESULT_OK) { Log.d(TAG, "Got Picture!"); Log.d(TAG,"File type - " + data.getType()); Uri photoUri = data.getData(); if (photoUri != null) { try { String[] filePathColumn = {MediaStore.Images.Media.DATA}; String[] orientationColumn = {MediaStore.Images.Media.ORIENTATION}; int orientation = -1; Cursor cursor = getContentResolver().query(photoUri, filePathColumn, null, null, null); cursor.moveToFirst(); int columnIndex = cursor.getColumnIndex(filePathColumn[0]); String filePath = cursor.getString(columnIndex); cursor.close(); cursor = getContentResolver().query(photoUri, orientationColumn, null, null, null); if(cursor != null && cursor.moveToFirst()){ orientation = cursor.getInt(cursor.getColumnIndex(orientationColumn[0])); } cursor.close(); HashMap<String, Integer> pRes = this.getImageResolutionSetting(); Bitmap shrunkenBitmap = FileUtilsHelper.shrinkBitmap(filePath, pRes.get("width"), pRes.get("height")); shrunkenBitmap = rotateBitmapToOrientation(shrunkenBitmap, orientation); String res = FileUtilsHelper.saveBitmapAsJpeg(shrunkenBitmap, this); Log.d(TAG,"File Path: " + res); shrunkenBitmap.recycle(); Intent editImage = new Intent(this, EditImage.class); editImage.addFlags(Intent.FLAG_ACTIVITY_FORWARD_RESULT); editImage.putExtra("stuff.path", res); startActivity(editImage); }catch(Exception e){ Toast.makeText(this, R.string.cant_save_image,Toast.LENGTH_SHORT).show(); } } } } break; } }} I don´t know what i´m doing wrong... I could really need a teacher on that :) Thx for your help dudes!!

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  • Automatically change div on mouseover and on timer

    - by IrishSaffa
    I'm a bit o a noob so any help would be great... What I need to do is have it so that the div associated to a specific li can change on hover as well as automatically change on a timer so that it scrolls through the option. here is my code: <script type="text/javascript"> $(function () { $("#switches li").mouseover(function () { var $this = $(this); $("#slides div").hide(); $("#slide" + $this.attr("id").replace(/switch/, "")).show(); }); }); </script> <div id="featured"> <ul id="switches"> <li id="switch1"><a href="activity_spa.html">Spa &amp; Wellness</a></li> <li id="switch2"><a href="#">Gala Venues</a></li> <li id="switch3"><a href="#">Dining</a></li> <li id="switch4"><a href="#">Shopping</a></li> <li id="switch5"><a href="#">Golf</a></li> <li id="switch6"><a href="#">Team Building</a></li> <li id="switch7"><a href="#">Equestrian</a></li> </ul> <div id="slides"> <div id="slide1"><img src="images/image2.jpg" alt="" /></div> <div id="slide2" style="display:none;"><img src="images/image6.jpg" alt="" /></div> <div id="slide3" style="display:none;"><img src="images/image1.jpg" alt="" /></div> <div id="slide4" style="display:none;"><img src="images/image3.jpg" alt="" /></div> <div id="slide5" style="display:none;"><img src="images/image5.jpg" alt="" /></div> <div id="slide6" style="display:none;"><img src="images/image7.jpg" alt="" /></div> <div id="slide7" style="display:none;"><img src="images/image4.jpg" alt="" /></div> </div> </div>

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  • [c#] SoundPlayer.PlaySync stopping prematurely

    - by JeffE
    I want to play a wav file synchronously on the gui thread, but my call to PlaySync is returning early (and prematurely stopping playback). The wav file is 2-3 minutes. Here's what my code looks like: //in gui code (event handler) //play first audio file JE_SP.playSound("example1.wav"); //do a few other statements doSomethingUnrelated(); //play another audio file JE_SP.playSound("example2.wav"); //library method written by me, called in gui code, but located in another assembly public static int playSound(string wavFile, bool synchronous = true, bool debug = true, string logFile = "", int loadTimeout = FIVE_MINUTES_IN_MS) { SoundPlayer sp = new SoundPlayer(); sp.LoadTimeout = loadTimeout; sp.SoundLocation = wavFile; sp.Load(); switch (synchronous) { case true: sp.PlaySync(); break; case false: sp.Play(); break; } if (debug) { string writeMe = "JE_SP: \r\n\tSoundLocation = " + sp.SoundLocation + "\r\n\t" + "Synchronous = " + synchronous.ToString(); JE_Log.logMessage(writeMe); } sp.Dispose(); sp = null; return 0; } Some things I've thought of are the load timeout, and playing the audio on another thread and then manually 'freeze' the gui by forcing the gui thread to wait for the duration of the sound file. I tried lengthening the load timeout, but that did nothing. I'm not quite sure what the best way to get the duration of a wav file is without using code written by somebody who isn't me/Microsoft. I suppose this can be calculated since I know the file size, and all of the encoding properties (bitrate, sample rate, sample size, etc) are consistent across all files I intend to play. Can somebody elaborate on how to calculate the duration of a wav file using this info? That is, if nobody has an idea about why PlaySync is returning early. Of Note: I encountered a similar problem in VB 6 a while ago, but that was caused by a timeout, which I don't suspect to be a problem here. Shorter (< 1min) files seem to play fine, so I might decide to manually edit the longer files down, then play them separately with multiple calls.

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  • How to make area outside of custom dialog view unclickable?

    - by portfoliobuilder
    I created a custom dialog (no, this is not dialog object) from an image and some other views. The conflict I am having with this custom dialog (again, this is a layout) is that the area around it closes the custom dialog. Is there a way I can make the outside area unclickable? I have tried wrapping the dialog view with a fullscreen frameLayout w/ transparent background, and then programmatically I set the frame attribute to setClickable(false). framelayout.setClickable(false); This does nothing. It still closes the dialog. Any other suggestions? Thank you in advance. This is my code: //used to disable background from closing the custom dialog private FrameLayout fl; @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); requestWindowFeature(Window.FEATURE_NO_TITLE); setContentView(R.layout.layout_dialog); btnContinue = (Button) findViewById(R.id.btnContinue); btnContinue.setOnClickListener(this); fl.setClickable(false); //background suppose to lock } @Override public void onClick(View v) { // TODO Auto-generated method stub switch (v.getId()) { case R.id.Continue: finish(); } break; } } I also have another class for broadcastReceiver public class DialogManagerBroadcastReceiver extends BroadcastReceiver { @Override public void onReceive(Context context, Intent intent) { if(IdeaPlayInterfaceApplication.isActivityVisible()){ Intent i=new Intent(context,CustomDialogActivity.class); i.setFlags(Intent.FLAG_ACTIVITY_NEW_TASK); context.startActivity(i); } } } The idea is that this custom dialog is not called at a specific instance, it is called every set amount of time no matter what I am doing in the application. I use an Intent and PendingIntent to repeatedly call this custom dialog over time. With something like this: cancelAlarmNotificationMonitoring(context); Calendar calendar = Calendar.getInstance(); Intent intent = new Intent(context, AlarmManagerBroadcastReceiver.class); PendingIntent pintent = PendingIntent.getBroadcast(context, 0, intent, 0); AlarmManager alarm = (AlarmManager) context.getSystemService(Context.ALARM_SERVICE); alarm.setRepeating(AlarmManager.RTC_WAKEUP,calendar.getTimeInMillis()+ALARM_INTERVAL,ALARM_INTERVAL, pintent); Hopefully this is more clear now.

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  • making a queue program

    - by seventhief
    Hi can someone help me making a queue program. i want to set the array[0] to be array[1] just in display but in real i am adding value at array[0]. i got how to run the add function to it. but i can't do the view and delete command that will view from ex. array[0] to array[4], when displayed array[1] to array[5] with the value inserted. #include <stdio.h> #include <stdlib.h> #define p printf #define s scanf int rear = 0; int front = 0; int *q_array = NULL; int size = 0; main() { int num, opt; char cont[] = { 'y' }; clrscr(); p("Queue Program\n\n"); p("Queue size: "); s("%d", &size); p("\n"); if(size > 0) { q_array = malloc(size * sizeof(int)); if(q_array == NULL) { p("ERROR: malloc() failed\n"); exit(2); } } else { p("ERROR: size should be positive integer\n"); exit(1); } while((cont[0] == 'y') || (cont[0] == 'Y')) { clrscr(); p("Queue Program"); p("\n\nQueue size: %d\n\n", size); p("MAIN MENU\n1. Add\n2. Delete\n3. View"); p("\n\nYour choice: "); s("%d", &opt); p("\n"); switch(opt) { case 1: if(rear==size) { p("You can't add more data"); } else { p("Enter data for Queue[%d]: ", rear+1); s("%d", &num); add(num); } break; case 2: delt(); break; case 3: view(); break; } p("\n\nDo you want to continue? (Y\/N)"); s("%s", &cont[0]); } } add(int a) { q_array[rear]=a; rear++; } delt() { if(front==rear) { p("Queue Empty"); } else { p("Queue[%d] = %d removed.", front, q_array[front]); front++; } } view() { int i; for(i=front;i<=rear;i++) p("\nQueue[%d] = %d", i, q_array[i]); }

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  • XMLNodes being appended to an XMLNode are "undefined"? Actionscript 2.0 is being unkind

    - by DigitalMercenary
    If anyone can offer an explanation for this one, I'd LOVE to see it! I was required to append a legacy application to display 20 random questions from an XML data source, as opposed to the total of 70 questions that are part of the original XML. No big deal, right? WRONG! I got it to work just fine in the end, but it's a total HACK! For some reason, some of the nodes that I am appending to a dynamically generated XML document are being returned as "undefined". I kept getting between 16 and 20 questions to render until I modified my iteration from a 'for' loop to a 'do while' loop with the appropriate number of XMLNodes as the condition of the 'do while' loop. Can anyone offer an explanation? Below is the code, with some notes for the reader : function editXML(xml:XML):XML { var node:XMLNode = xml.firstChild; var newNode:XMLNode = new XMLNode(); var nodeArray:Array = new Array(); var usedNodes:Array = new Array(); var totalNodes:Number = node.lastChild.childNodes.length - 1; var nextNode:Number; var returnNode:XMLNode = new XMLNode(); var tempNode:XMLNode; var buildNode:XMLNode; var addNode:Boolean = true; var tempXML:XML = new XML(); var pagesNode:XMLNode = tempXML.createElement("pages"); tempXML.appendChild(pagesNode); tempXML.appendChild(node.childNodes[0]); tempXML.appendChild(node.childNodes[1]); tempXML.appendChild(node.childNodes[2]); var questionsNode:XMLNode = tempXML.createElement("pages"); tempXML.firstChild.appendChild(questionsNode); do { nextNode = Math.floor(Math.random()*totalNodes); **//random number to represent random node** //trace(nextNode + " nextNode"); **//check usedNodes Array to look for node.childNodes[nextNode]. If it already exists, skip and reloop.** trace(node.childNodes[1].childNodes[nextNode] + " : pre building Node " + totalNodes); if(usedNodes.length == 0) { buildNode = new XMLNode(); buildNode.nodeName = node.childNodes[1].childNodes[nextNode].nodeName; buildNode.nodeValue = node.childNodes[1].childNodes[nextNode].nodeValue; tempXML.firstChild.lastChild.appendChild(node.childNodes[1].childNodes[nextNode]) usedNodes.push(node.childNodes[1].childNodes[nextNode]); nodeArray.push(node.childNodes[1].childNodes[nextNode]); trace("adding first node : " + nodeArray.length); addNode = false; } else { for(var j:Number = 0; j < usedNodes.length; j++) { if(usedNodes[j] == node.childNodes[1].childNodes[nextNode]) { addNode = false; trace("skipping node : " + nodeArray.length); } } } **//if node not in usedNodes, add node to XML** if(addNode) { trace(node.childNodes[1].childNodes[nextNode] + " : building Node"); **//This trace statement produced a valid node** tempXML.firstChild.lastChild.appendChild(node.childNodes[1].childNodes[nextNode]); **//Before modifying the code from adding nodes to the xml from an Array called 'nodeArray' in a for loop to adding nodes directly to the xml in a do while loop with the length of the xml node used to retrieve data for the questions as the condition, I was not always getting 20 questions. Some of the nodes were being rendered as 'undefined' and not appended to the xml, even though they were traced and proven valid before the attemp to append them to the xml was made** usedNodes.push(node.childNodes[1].childNodes[nextNode]); } addNode = true; } while(tempXML.firstChild.lastChild.childNodes.length <= 19); trace(tempXML.firstChild.lastChild.childNodes.length + " final nodes Length"); courseXML = tempXML; //removes the old question list of 70 and replaces it with the new question list of 20. Question list is the last node. return tempXML; } If I had my choice, I would have rebuilt the whole application in Flex with AS3. I didn't have that choice. If anyone can explain this mystery, PLEASE DO! Thank you in advance!

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  • database design help for game / user levels / progress

    - by sprugman
    Sorry this got long and all prose-y. I'm creating my first truly gamified web app and could use some help thinking about how to structure the data. The Set-up Users need to accomplish tasks in each of several categories before they can move up a level. I've got my Users, Tasks, and Categories tables, and a UserTasks table which joins the three. ("User 3 has added Task 42 in Category 8. Now they've completed it.") That's all fine and working wonderfully. The Challenge I'm not sure of the best way to track the progress in the individual categories toward each level. The "business" rules are: You have to achieve a certain number of points in each category to move up. If you get the number of points needed in Cat 8, but still have other work to do to complete the level, any new Cat 8 points count toward your overall score, but don't "roll over" into the next level. The number of Categories is small (five currently) and unlikely to change often, but by no means absolutely fixed. The number of points needed to level-up will vary per level, probably by a formula, or perhaps a lookup table. So the challenge is to track each user's progress toward the next level in each category. I've thought of a few potential approaches: Possible Solutions Add a column to the users table for each category and reset them all to zero each time a user levels-up. Have a separate UserProgress table with a row for each category for each user and the number of points they have. (Basically a Many-to-Many version of #1.) Add a userLevel column to the UserTasks table and use that to derive their progress with some kind of SUM statement. Their current level will be a simple int in the User table. Pros & Cons (1) seems like by far the most straightforward, but it's also the least flexible. Perhaps I could use a naming convention based on the category ids to help overcome some of that. (With code like "select cats; for each cat, get the value from Users.progress_{cat.id}.") It's also the one where I lose the most data -- I won't know which points counted toward leveling up. I don't have a need in mind for that, so maybe I don't care about that. (2) seems complicated: every time I add or subtract a user or a category, I have to maintain the other table. I foresee synchronization challenges. (3) Is somewhere in between -- cleaner than #2, but less intuitive than #1. In order to find out where a user is, I'd have mildly complex SQL like: SELECT categoryId, SUM(points) from UserTasks WHERE userId={user.id} & countsTowardLevel={user.level} groupBy categoryId Hmm... that doesn't seem so bad. I think I'm talking myself into #3 here, but would love any input, advice or other ideas.

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  • Referencing variables in a structure / C++

    - by user1628622
    Below, I provided a minimal example of code I created. I managed to get this code working, but I'm not sure if the practice being employed is sound. In essence, what I am trying to do is have the 'Parameter' class reference select elements in the 'States' class, so variables in States can be changed via Parameters. Questions I have: is the approach taken OK? If not, is there a better way to achieve what I am aiming for? Example code: struct VAR_TYPE{ public: bool is_fixed; // If is_fixed = true, then variable is a parameter double value; // Numerical value std::string name; // Description of variable (to identify it by name) }; struct NODE{ public: VAR_TYPE X, Y, Z; /* VAR_TYPE is a structure of primitive types */ }; class States{ private: std::vector <NODE_ptr> node; // shared ptr to struct NODE std::vector <PROP_DICTIONARY_ptr> property; // CAN NOT be part of Parameter std::vector <ELEMENT_ptr> element; // CAN NOT be part of Parameter public: /* ect */ void set_X_reference ( Parameter &T , int i ) { T.push_var( &node[i]->X ); } void set_Y_reference ( Parameter &T , int i ) { T.push_var( &node[i]->Y ); } void set_Z_reference ( Parameter &T , int i ) { T.push_var( &node[i]->Z ); } bool get_node_bool_X( int i ) { return node[i]->X.is_fixed; } // repeat for Y and Z }; class Parameter{ private: std::vector <VAR_TYPE*> var; public: /* ect */ }; int main(){ States S; Parameter P; /* Here I initialize and set S, and do other stuff */ // Now I assign components in States to Parameters for(int n=0 ; n<S.size_of_nodes() ; n++ ){ if ( S.get_node_bool_X(n)==true ){ S.set_X_reference ( P , n ); }; // repeat if statement for Y and Z }; /* Now P points selected to data in S, and I can * modify the contents of S through P */ return 0; }; Update The reason this issue cropped up is I am working with Fortran legacy code. To sum up this Fotran code - it's a numerical simulation of a flight vehicle. This code has a fairly rigid procedural framework one must work within, which comes with a pre-defined list of allowable Fortran types. The Fortran glue code can create an instance of a C++ object (in actuality, a reference from the perspective of Fortran), but is not aware what is contained in it (other means are used to extract C++ data into Fortran). The problem that I encountered is when a C++ module is dynamically linked to the Fortran glue code, C++ objects have to be initialized each instance the C++ code is called. This happens by virtue of how the Fortran template is defined. To avoid this cycle of re-initializing objects, I plan to use 'State' as a container class. The Fortran code allows a 'State' object, which has an arbitrary definition; but I plan to use it to harness all relevant information about the model. The idea is to use the Parameters class (which is exposed and updated by the Fortran code) to update variables in States.

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  • I want to check if $('#td1').text() === "x"?

    - by M.z
    I want to check if innerHtml have X or O , so i can not add again any thing else , but it's not working . it stop after adding the check code , I'm trying here to do a simple X O game to get more familiar with javascript and jquery . also I'm not sure if can do this with jQuery . <script type="text/javascript" > function ranFun() { return Math.floor((Math.random() * 9) + 1); } var a; function Elment(a) { document.getElementById("td" + a).innerHTML = "O"; } function call() { var x = ranFun(); switch (x) { case 1:case 2 :case 3: case 4 :case 5 : case 6 : case 7 : case 8 : case 9 : Elment(x); break; default: break; } } function tdElm(c) { if ($('#td1').text() === "x" || $('#td1').text() == "o") return false; else { document.getElementById("td" + c).innerHTML = "x"; call(); } } </script> <BODY> <center> <h1 >" X ,O Game "</h1> <table > <tr> <td id="td1" onclick="tdElm(1);" ></td> <td id="td2" onclick="tdElm(2);"></td> <td id="td3" onclick="tdElm(3);"></td> </tr> <tr> <td id="td4" onclick="tdElm(4);"></td> <td id="td5" onclick="tdElm(5);"></td> <td id="td6" onclick="tdElm(6);"></td> </tr> <tr> <td id="td7" onclick="tdElm(7);"></td> <td id="td8" onclick="tdElm(8);"></td> <td id="td9" onclick="tdElm(9);"></td> </tr> </table> </center> </BODY>

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  • Dependency Injection in ASP.NET MVC NerdDinner App using Unity 2.0

    - by shiju
    In my previous post Dependency Injection in ASP.NET MVC NerdDinner App using Ninject, we did dependency injection in NerdDinner application using Ninject. In this post, I demonstrate how to apply Dependency Injection in ASP.NET MVC NerdDinner App using Microsoft Unity Application Block (Unity) v 2.0.Unity 2.0Unity 2.0 is available on Codeplex at http://unity.codeplex.com . In earlier versions of Unity, the ObjectBuilder generic dependency injection mechanism, was distributed as a separate assembly, is now integrated with Unity core assembly. So you no longer need to reference the ObjectBuilder assembly in your applications. Two additional Built-In Lifetime Managers - HierarchicalifetimeManager and PerResolveLifetimeManager have been added to Unity 2.0.Dependency Injection in NerdDinner using UnityIn my Ninject post on NerdDinner, we have discussed the interfaces and concrete types of NerdDinner application and how to inject dependencies controller constructors. The following steps will configure Unity 2.0 to apply controller injection in NerdDinner application. Step 1 – Add reference for Unity Application BlockOpen the NerdDinner solution and add  reference to Microsoft.Practices.Unity.dll and Microsoft.Practices.Unity.Configuration.dllYou can download Unity from at http://unity.codeplex.com .Step 2 – Controller Factory for Unity The controller factory is responsible for creating controller instances.We extend the built in default controller factory with our own factory for working Unity with ASP.NET MVC. public class UnityControllerFactory : DefaultControllerFactory {     protected override IController GetControllerInstance(RequestContext reqContext, Type controllerType)     {         IController controller;         if (controllerType == null)             throw new HttpException(                     404, String.Format(                         "The controller for path '{0}' could not be found" +         "or it does not implement IController.",                     reqContext.HttpContext.Request.Path));           if (!typeof(IController).IsAssignableFrom(controllerType))             throw new ArgumentException(                     string.Format(                         "Type requested is not a controller: {0}",                         controllerType.Name),                         "controllerType");         try         {             controller = MvcUnityContainer.Container.Resolve(controllerType)                             as IController;         }         catch (Exception ex)         {             throw new InvalidOperationException(String.Format(                                     "Error resolving controller {0}",                                     controllerType.Name), ex);         }         return controller;     }   }   public static class MvcUnityContainer {     public static IUnityContainer Container { get; set; } }  Step 3 – Register Types and Set Controller Factory private void ConfigureUnity() {     //Create UnityContainer               IUnityContainer container = new UnityContainer()     .RegisterType<IFormsAuthentication, FormsAuthenticationService>()     .RegisterType<IMembershipService, AccountMembershipService>()     .RegisterInstance<MembershipProvider>(Membership.Provider)     .RegisterType<IDinnerRepository, DinnerRepository>();     //Set container for Controller Factory     MvcUnityContainer.Container = container;     //Set Controller Factory as UnityControllerFactory     ControllerBuilder.Current.SetControllerFactory(                         typeof(UnityControllerFactory));            } Unity 2.0 provides a fluent interface for type configuration. Now you can call all the methods in a single statement.The above Unity configuration specified in the ConfigureUnity method tells that, to inject instance of DinnerRepositiry when there is a request for IDinnerRepositiry and  inject instance of FormsAuthenticationService when there is a request for IFormsAuthentication and inject instance of AccountMembershipService when there is a request for IMembershipService. The AccountMembershipService class has a dependency with ASP.NET Membership provider. So we configure that inject the instance of Membership Provider.After the registering the types, we set UnityControllerFactory as the current controller factory. //Set container for Controller Factory MvcUnityContainer.Container = container; //Set Controller Factory as UnityControllerFactory ControllerBuilder.Current.SetControllerFactory(                     typeof(UnityControllerFactory)); When you register a type  by using the RegisterType method, the default behavior is for the container to use a transient lifetime manager. It creates a new instance of the registered, mapped, or requested type each time you call the Resolve or ResolveAll method or when the dependency mechanism injects instances into other classes. The following are the LifetimeManagers provided by Unity 2.0ContainerControlledLifetimeManager - Implements a singleton behavior for objects. The object is disposed of when you dispose of the container.ExternallyControlledLifetimeManager - Implements a singleton behavior but the container doesn't hold a reference to object which will be disposed of when out of scope.HierarchicalifetimeManager - Implements a singleton behavior for objects. However, child containers don't share instances with parents.PerResolveLifetimeManager - Implements a behavior similar to the transient lifetime manager except that instances are reused across build-ups of the object graph.PerThreadLifetimeManager - Implements a singleton behavior for objects but limited to the current thread.TransientLifetimeManager - Returns a new instance of the requested type for each call. (default behavior)We can also create custome lifetime manager for Unity container. The following code creating a custom lifetime manager to store container in the current HttpContext. public class HttpContextLifetimeManager<T> : LifetimeManager, IDisposable {     public override object GetValue()     {         return HttpContext.Current.Items[typeof(T).AssemblyQualifiedName];     }     public override void RemoveValue()     {         HttpContext.Current.Items.Remove(typeof(T).AssemblyQualifiedName);     }     public override void SetValue(object newValue)     {         HttpContext.Current.Items[typeof(T).AssemblyQualifiedName]             = newValue;     }     public void Dispose()     {         RemoveValue();     } }  Step 4 – Modify Global.asax.cs for configure Unity container In the Application_Start event, we call the ConfigureUnity method for configuring the Unity container and set controller factory as UnityControllerFactory void Application_Start() {     RegisterRoutes(RouteTable.Routes);       ViewEngines.Engines.Clear();     ViewEngines.Engines.Add(new MobileCapableWebFormViewEngine());     ConfigureUnity(); }Download CodeYou can download the modified NerdDinner code from http://nerddinneraddons.codeplex.com

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  • Parallelism in .NET – Part 2, Simple Imperative Data Parallelism

    - by Reed
    In my discussion of Decomposition of the problem space, I mentioned that Data Decomposition is often the simplest abstraction to use when trying to parallelize a routine.  If a problem can be decomposed based off the data, we will often want to use what MSDN refers to as Data Parallelism as our strategy for implementing our routine.  The Task Parallel Library in .NET 4 makes implementing Data Parallelism, for most cases, very simple. Data Parallelism is the main technique we use to parallelize a routine which can be decomposed based off data.  Data Parallelism refers to taking a single collection of data, and having a single operation be performed concurrently on elements in the collection.  One side note here: Data Parallelism is also sometimes referred to as the Loop Parallelism Pattern or Loop-level Parallelism.  In general, for this series, I will try to use the terminology used in the MSDN Documentation for the Task Parallel Library.  This should make it easier to investigate these topics in more detail. Once we’ve determined we have a problem that, potentially, can be decomposed based on data, implementation using Data Parallelism in the TPL is quite simple.  Let’s take our example from the Data Decomposition discussion – a simple contrast stretching filter.  Here, we have a collection of data (pixels), and we need to run a simple operation on each element of the pixel.  Once we know the minimum and maximum values, we most likely would have some simple code like the following: for (int row=0; row < pixelData.GetUpperBound(0); ++row) { for (int col=0; col < pixelData.GetUpperBound(1); ++col) { pixelData[row, col] = AdjustContrast(pixelData[row, col], minPixel, maxPixel); } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } This simple routine loops through a two dimensional array of pixelData, and calls the AdjustContrast routine on each pixel. As I mentioned, when you’re decomposing a problem space, most iteration statements are potentially candidates for data decomposition.  Here, we’re using two for loops – one looping through rows in the image, and a second nested loop iterating through the columns.  We then perform one, independent operation on each element based on those loop positions. This is a prime candidate – we have no shared data, no dependencies on anything but the pixel which we want to change.  Since we’re using a for loop, we can easily parallelize this using the Parallel.For method in the TPL: Parallel.For(0, pixelData.GetUpperBound(0), row => { for (int col=0; col < pixelData.GetUpperBound(1); ++col) { pixelData[row, col] = AdjustContrast(pixelData[row, col], minPixel, maxPixel); } }); Here, by simply changing our first for loop to a call to Parallel.For, we can parallelize this portion of our routine.  Parallel.For works, as do many methods in the TPL, by creating a delegate and using it as an argument to a method.  In this case, our for loop iteration block becomes a delegate creating via a lambda expression.  This lets you write code that, superficially, looks similar to the familiar for loop, but functions quite differently at runtime. We could easily do this to our second for loop as well, but that may not be a good idea.  There is a balance to be struck when writing parallel code.  We want to have enough work items to keep all of our processors busy, but the more we partition our data, the more overhead we introduce.  In this case, we have an image of data – most likely hundreds of pixels in both dimensions.  By just parallelizing our first loop, each row of pixels can be run as a single task.  With hundreds of rows of data, we are providing fine enough granularity to keep all of our processors busy. If we parallelize both loops, we’re potentially creating millions of independent tasks.  This introduces extra overhead with no extra gain, and will actually reduce our overall performance.  This leads to my first guideline when writing parallel code: Partition your problem into enough tasks to keep each processor busy throughout the operation, but not more than necessary to keep each processor busy. Also note that I parallelized the outer loop.  I could have just as easily partitioned the inner loop.  However, partitioning the inner loop would have led to many more discrete work items, each with a smaller amount of work (operate on one pixel instead of one row of pixels).  My second guideline when writing parallel code reflects this: Partition your problem in a way to place the most work possible into each task. This typically means, in practice, that you will want to parallelize the routine at the “highest” point possible in the routine, typically the outermost loop.  If you’re looking at parallelizing methods which call other methods, you’ll want to try to partition your work high up in the stack – as you get into lower level methods, the performance impact of parallelizing your routines may not overcome the overhead introduced. Parallel.For works great for situations where we know the number of elements we’re going to process in advance.  If we’re iterating through an IList<T> or an array, this is a typical approach.  However, there are other iteration statements common in C#.  In many situations, we’ll use foreach instead of a for loop.  This can be more understandable and easier to read, but also has the advantage of working with collections which only implement IEnumerable<T>, where we do not know the number of elements involved in advance. As an example, lets take the following situation.  Say we have a collection of Customers, and we want to iterate through each customer, check some information about the customer, and if a certain case is met, send an email to the customer and update our instance to reflect this change.  Normally, this might look something like: foreach(var customer in customers) { // Run some process that takes some time... DateTime lastContact = theStore.GetLastContact(customer); TimeSpan timeSinceContact = DateTime.Now - lastContact; // If it's been more than two weeks, send an email, and update... if (timeSinceContact.Days > 14) { theStore.EmailCustomer(customer); customer.LastEmailContact = DateTime.Now; } } Here, we’re doing a fair amount of work for each customer in our collection, but we don’t know how many customers exist.  If we assume that theStore.GetLastContact(customer) and theStore.EmailCustomer(customer) are both side-effect free, thread safe operations, we could parallelize this using Parallel.ForEach: Parallel.ForEach(customers, customer => { // Run some process that takes some time... DateTime lastContact = theStore.GetLastContact(customer); TimeSpan timeSinceContact = DateTime.Now - lastContact; // If it's been more than two weeks, send an email, and update... if (timeSinceContact.Days > 14) { theStore.EmailCustomer(customer); customer.LastEmailContact = DateTime.Now; } }); Just like Parallel.For, we rework our loop into a method call accepting a delegate created via a lambda expression.  This keeps our new code very similar to our original iteration statement, however, this will now execute in parallel.  The same guidelines apply with Parallel.ForEach as with Parallel.For. The other iteration statements, do and while, do not have direct equivalents in the Task Parallel Library.  These, however, are very easy to implement using Parallel.ForEach and the yield keyword. Most applications can benefit from implementing some form of Data Parallelism.  Iterating through collections and performing “work” is a very common pattern in nearly every application.  When the problem can be decomposed by data, we often can parallelize the workload by merely changing foreach statements to Parallel.ForEach method calls, and for loops to Parallel.For method calls.  Any time your program operates on a collection, and does a set of work on each item in the collection where that work is not dependent on other information, you very likely have an opportunity to parallelize your routine.

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  • Parallelism in .NET – Part 4, Imperative Data Parallelism: Aggregation

    - by Reed
    In the article on simple data parallelism, I described how to perform an operation on an entire collection of elements in parallel.  Often, this is not adequate, as the parallel operation is going to be performing some form of aggregation. Simple examples of this might include taking the sum of the results of processing a function on each element in the collection, or finding the minimum of the collection given some criteria.  This can be done using the techniques described in simple data parallelism, however, special care needs to be taken into account to synchronize the shared data appropriately.  The Task Parallel Library has tools to assist in this synchronization. The main issue with aggregation when parallelizing a routine is that you need to handle synchronization of data.  Since multiple threads will need to write to a shared portion of data.  Suppose, for example, that we wanted to parallelize a simple loop that looked for the minimum value within a dataset: double min = double.MaxValue; foreach(var item in collection) { double value = item.PerformComputation(); min = System.Math.Min(min, value); } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } This seems like a good candidate for parallelization, but there is a problem here.  If we just wrap this into a call to Parallel.ForEach, we’ll introduce a critical race condition, and get the wrong answer.  Let’s look at what happens here: // Buggy code! Do not use! double min = double.MaxValue; Parallel.ForEach(collection, item => { double value = item.PerformComputation(); min = System.Math.Min(min, value); }); This code has a fatal flaw: min will be checked, then set, by multiple threads simultaneously.  Two threads may perform the check at the same time, and set the wrong value for min.  Say we get a value of 1 in thread 1, and a value of 2 in thread 2, and these two elements are the first two to run.  If both hit the min check line at the same time, both will determine that min should change, to 1 and 2 respectively.  If element 1 happens to set the variable first, then element 2 sets the min variable, we’ll detect a min value of 2 instead of 1.  This can lead to wrong answers. Unfortunately, fixing this, with the Parallel.ForEach call we’re using, would require adding locking.  We would need to rewrite this like: // Safe, but slow double min = double.MaxValue; // Make a "lock" object object syncObject = new object(); Parallel.ForEach(collection, item => { double value = item.PerformComputation(); lock(syncObject) min = System.Math.Min(min, value); }); This will potentially add a huge amount of overhead to our calculation.  Since we can potentially block while waiting on the lock for every single iteration, we will most likely slow this down to where it is actually quite a bit slower than our serial implementation.  The problem is the lock statement – any time you use lock(object), you’re almost assuring reduced performance in a parallel situation.  This leads to two observations I’ll make: When parallelizing a routine, try to avoid locks. That being said: Always add any and all required synchronization to avoid race conditions. These two observations tend to be opposing forces – we often need to synchronize our algorithms, but we also want to avoid the synchronization when possible.  Looking at our routine, there is no way to directly avoid this lock, since each element is potentially being run on a separate thread, and this lock is necessary in order for our routine to function correctly every time. However, this isn’t the only way to design this routine to implement this algorithm.  Realize that, although our collection may have thousands or even millions of elements, we have a limited number of Processing Elements (PE).  Processing Element is the standard term for a hardware element which can process and execute instructions.  This typically is a core in your processor, but many modern systems have multiple hardware execution threads per core.  The Task Parallel Library will not execute the work for each item in the collection as a separate work item. Instead, when Parallel.ForEach executes, it will partition the collection into larger “chunks” which get processed on different threads via the ThreadPool.  This helps reduce the threading overhead, and help the overall speed.  In general, the Parallel class will only use one thread per PE in the system. Given the fact that there are typically fewer threads than work items, we can rethink our algorithm design.  We can parallelize our algorithm more effectively by approaching it differently.  Because the basic aggregation we are doing here (Min) is communitive, we do not need to perform this in a given order.  We knew this to be true already – otherwise, we wouldn’t have been able to parallelize this routine in the first place.  With this in mind, we can treat each thread’s work independently, allowing each thread to serially process many elements with no locking, then, after all the threads are complete, “merge” together the results. This can be accomplished via a different set of overloads in the Parallel class: Parallel.ForEach<TSource,TLocal>.  The idea behind these overloads is to allow each thread to begin by initializing some local state (TLocal).  The thread will then process an entire set of items in the source collection, providing that state to the delegate which processes an individual item.  Finally, at the end, a separate delegate is run which allows you to handle merging that local state into your final results. To rewriting our routine using Parallel.ForEach<TSource,TLocal>, we need to provide three delegates instead of one.  The most basic version of this function is declared as: public static ParallelLoopResult ForEach<TSource, TLocal>( IEnumerable<TSource> source, Func<TLocal> localInit, Func<TSource, ParallelLoopState, TLocal, TLocal> body, Action<TLocal> localFinally ) The first delegate (the localInit argument) is defined as Func<TLocal>.  This delegate initializes our local state.  It should return some object we can use to track the results of a single thread’s operations. The second delegate (the body argument) is where our main processing occurs, although now, instead of being an Action<T>, we actually provide a Func<TSource, ParallelLoopState, TLocal, TLocal> delegate.  This delegate will receive three arguments: our original element from the collection (TSource), a ParallelLoopState which we can use for early termination, and the instance of our local state we created (TLocal).  It should do whatever processing you wish to occur per element, then return the value of the local state after processing is completed. The third delegate (the localFinally argument) is defined as Action<TLocal>.  This delegate is passed our local state after it’s been processed by all of the elements this thread will handle.  This is where you can merge your final results together.  This may require synchronization, but now, instead of synchronizing once per element (potentially millions of times), you’ll only have to synchronize once per thread, which is an ideal situation. Now that I’ve explained how this works, lets look at the code: // Safe, and fast! double min = double.MaxValue; // Make a "lock" object object syncObject = new object(); Parallel.ForEach( collection, // First, we provide a local state initialization delegate. () => double.MaxValue, // Next, we supply the body, which takes the original item, loop state, // and local state, and returns a new local state (item, loopState, localState) => { double value = item.PerformComputation(); return System.Math.Min(localState, value); }, // Finally, we provide an Action<TLocal>, to "merge" results together localState => { // This requires locking, but it's only once per used thread lock(syncObj) min = System.Math.Min(min, localState); } ); Although this is a bit more complicated than the previous version, it is now both thread-safe, and has minimal locking.  This same approach can be used by Parallel.For, although now, it’s Parallel.For<TLocal>.  When working with Parallel.For<TLocal>, you use the same triplet of delegates, with the same purpose and results. Also, many times, you can completely avoid locking by using a method of the Interlocked class to perform the final aggregation in an atomic operation.  The MSDN example demonstrating this same technique using Parallel.For uses the Interlocked class instead of a lock, since they are doing a sum operation on a long variable, which is possible via Interlocked.Add. By taking advantage of local state, we can use the Parallel class methods to parallelize algorithms such as aggregation, which, at first, may seem like poor candidates for parallelization.  Doing so requires careful consideration, and often requires a slight redesign of the algorithm, but the performance gains can be significant if handled in a way to avoid excessive synchronization.

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  • Visual Studio 2010 Extension Manager (and the new VS 2010 PowerCommands Extension)

    - by ScottGu
    This is the twenty-third in a series of blog posts I’m doing on the VS 2010 and .NET 4 release. Today’s blog post covers some of the extensibility improvements made in VS 2010 – as well as a cool new "PowerCommands for Visual Studio 2010” extension that Microsoft just released (and which can be downloaded and used for free). [In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu] Extensibility in VS 2010 VS 2010 provides a much richer extensibility model than previous releases.  Anyone can build extensions that add, customize, and light-up the Visual Studio 2010 IDE, Code Editors, Project System and associated Designers. VS 2010 Extensions can be created using the new MEF (Managed Extensibility Framework) which is built-into .NET 4.  You can learn more about how to create VS 2010 extensions from this this blog post from the Visual Studio Team Blog. VS 2010 Extension Manager Developers building extensions can distribute them on their own (via their own web-sites or by selling them).  Visual Studio 2010 also now includes a built-in “Extension Manager” within the IDE that makes it much easier for developers to find, download, and enable extensions online.  You can launch the “Extension Manager” by selecting the Tools->Extension Manager menu option: This loads an “Extension Manager” dialog which accesses an “online gallery” at Microsoft, and then populates a list of available extensions that you can optionally download and enable within your copy of Visual Studio: There are already hundreds of cool extensions populated within the online gallery.  You can browse them by category (use the tree-view on the top-left to filter them).  Clicking “download” on any of the extensions will download, install, and enable it. PowerCommands for Visual Studio 2010 This weekend Microsoft released the free PowerCommands for Visual Studio 2010 extension to the online gallery.  You can learn more about it here, and download and install it via the “Extension Manager” above (search for PowerCommands to find it). The PowerCommands download adds dozens of useful commands to Visual Studio 2010.  Below is a screen-shot of just a few of the useful commands that it adds to the Solution Explorer context menus: Below is a list of all the commands included with this weekend’s PowerCommands for Visual Studio 2010 release: Enable/Disable PowerCommands in Options dialog This feature allows you to select which commands to enable in the Visual Studio IDE. Point to the Tools menu, then click Options. Expand the PowerCommands options, then click Commands. Check the commands you would like to enable. Note: All power commands are initially defaulted Enabled. Format document on save / Remove and Sort Usings on save The Format document on save option formats the tabs, spaces, and so on of the document being saved. It is equivalent to pointing to the Edit menu, clicking Advanced, and then clicking Format Document. The Remove and sort usings option removes unused using statements and sorts the remaining using statements in the document being saved. Note: The Remove and sort usings option is only available for C# documents. Format document on save and Remove and sort usings both are initially defaulted OFF. Clear All Panes This command clears all output panes. It can be executed from the button on the toolbar of the Output window. Copy Path This command copies the full path of the currently selected item to the clipboard. It can be executed by right-clicking one of these nodes in the Solution Explorer: The solution node; A project node; Any project item node; Any folder. Email CodeSnippet To email the lines of text you select in the code editor, right-click anywhere in the editor and then click Email CodeSnippet. Insert Guid Attribute This command adds a Guid attribute to a selected class. From the code editor, right-click anywhere within the class definition, then click Insert Guid Attribute. Show All Files This command shows the hidden files in all projects displayed in the Solution Explorer when the solution node is selected. It enhances the Show All Files button, which normally shows only the hidden files in the selected project node. Undo Close This command reopens a closed document , returning the cursor to its last position. To reopen the most recently closed document, point to the Edit menu, then click Undo Close. Alternately, you can use the CtrlShiftZ shortcut. To reopen any other recently closed document, point to the View menu, click Other Windows, and then click Undo Close Window. The Undo Close window appears, typically next to the Output window. Double-click any document in the list to reopen it. Collapse Projects This command collapses a project or projects in the Solution Explorer starting from the root selected node. Collapsing a project can increase the readability of the solution. This command can be executed from three different places: solution, solution folders and project nodes respectively. Copy Class This command copies a selected class entire content to the clipboard, renaming the class. This command is normally followed by a Paste Class command, which renames the class to avoid a compilation error. It can be executed from a single project item or a project item with dependent sub items. Paste Class This command pastes a class entire content from the clipboard, renaming the class to avoid a compilation error. This command is normally preceded by a Copy Class command. It can be executed from a project or folder node. Copy References This command copies a reference or set of references to the clipboard. It can be executed from the references node, a single reference node or set of reference nodes. Paste References This command pastes a reference or set of references from the clipboard. It can be executed from different places depending on the type of project. For CSharp projects it can be executed from the references node. For Visual Basic and Website projects it can be executed from the project node. Copy As Project Reference This command copies a project as a project reference to the clipboard. It can be executed from a project node. Edit Project File This command opens the MSBuild project file for a selected project inside Visual Studio. It combines the existing Unload Project and Edit Project commands. Open Containing Folder This command opens a Windows Explorer window pointing to the physical path of a selected item. It can be executed from a project item node Open Command Prompt This command opens a Visual Studio command prompt pointing to the physical path of a selected item. It can be executed from four different places: solution, project, folder and project item nodes respectively. Unload Projects This command unloads all projects in a solution. This can be useful in MSBuild scenarios when multiple projects are being edited. This command can be executed from the solution node. Reload Projects This command reloads all unloaded projects in a solution. It can be executed from the solution node. Remove and Sort Usings This command removes and sort using statements for all classes given a project. It is useful, for example, in removing or organizing the using statements generated by a wizard. This command can be executed from a solution node or a single project node. Extract Constant This command creates a constant definition statement for a selected text. Extracting a constant effectively names a literal value, which can improve readability. This command can be executed from the code editor by right-clicking selected text. Clear Recent File List This command clears the Visual Studio recent file list. The Clear Recent File List command brings up a Clear File dialog which allows any or all recent files to be selected. Clear Recent Project List This command clears the Visual Studio recent project list. The Clear Recent Project List command brings up a Clear File dialog which allows any or all recent projects to be selected. Transform Templates This command executes a custom tool with associated text templates items. It can be executed from a DSL project node or a DSL folder node. Close All This command closes all documents. It can be executed from a document tab. How to temporarily disable extensions Extensions provide a great way to make Visual Studio even more powerful, and can help improve your overall productivity.  One thing to keep in mind, though, is that extensions run within the Visual Studio process (DevEnv.exe) and so a bug within an extension can impact both the stability and performance of Visual Studio.  If you ever run into a situation where things seem slower than they should, or if you crash repeatedly, please temporarily disable any installed extensions and see if that fixes the problem.  You can do this for extensions that were installed via the online gallery by re-running the extension manager (using the Tools->Extension Manager menu option) and by selecting the “Installed Extensions” node on the top-left of the dialog – and then by clicking “Disable” on any of the extensions within your installed list: Hope this helps, Scott

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  • Entity Framework version 1- Brief Synopsis and Tips &ndash; Part 1

    - by Rohit Gupta
    To Do Eager loading use Projections (for e.g. from c in context.Contacts select c, c.Addresses)  or use Include Query Builder Methods (Include(“Addresses”)) If there is multi-level hierarchical Data then to eager load all the relationships use Include Query Builder methods like customers.Include("Order.OrderDetail") to include Order and OrderDetail collections or use customers.Include("Order.OrderDetail.Location") to include all Order, OrderDetail and location collections with a single include statement =========================================================================== If the query uses Joins then Include() Query Builder method will be ignored, use Nested Queries instead If the query does projections then Include() Query Builder method will be ignored Use Address.ContactReference.Load() OR Contact.Addresses.Load() if you need to Deferred Load Specific Entity – This will result in extra round trips to the database ObjectQuery<> cannot return anonymous types... it will return a ObjectQuery<DBDataRecord> Only Include method can be added to Linq Query Methods Any Linq Query method can be added to Query Builder methods. If you need to append a Query Builder Method (other than Include) after a LINQ method  then cast the IQueryable<Contact> to ObjectQuery<Contact> and then append the Query Builder method to it =========================================================================== Query Builder methods are Select, Where, Include Methods which use Entity SQL as parameters e.g. "it.StartDate, it.EndDate" When Query Builder methods do projection then they return ObjectQuery<DBDataRecord>, thus to iterate over this collection use contact.Item[“Name”].ToString() When Linq To Entities methods do projection, they return collection of anonymous types --- thus the collection is strongly typed and supports Intellisense EF Object Context can track changes only on Entities, not on Anonymous types. If you use a Defining Query for a EntitySet then the EntitySet becomes readonly since a Defining Query is the same as a View (which is treated as a ReadOnly by default). However if you want to use this EntitySet for insert/update/deletes then we need to map stored procs (as created in the DB) to the insert/update/delete functions of the Entity in the Designer You can use either Execute method or ToList() method to bind data to datasources/bindingsources If you use the Execute Method then remember that you can traverse through the ObjectResult<> collection (returned by Execute) only ONCE. In WPF use ObservableCollection to bind to data sources , for keeping track of changes and letting EF send updates to the DB automatically. Use Extension Methods to add logic to Entities. For e.g. create extension methods for the EntityObject class. Create a method in ObjectContext Partial class and pass the entity as a parameter, then call this method as desired from within each entity. ================================================================ DefiningQueries and Stored Procedures: For Custom Entities, one can use DefiningQuery or Stored Procedures. Thus the Custom Entity Collection will be populated using the DefiningQuery (of the EntitySet) or the Sproc. If you use Sproc to populate the EntityCollection then the query execution is immediate and this execution happens on the Server side and any filters applied will be applied in the Client App. If we use a DefiningQuery then these queries are composable, meaning the filters (if applied to the entityset) will all be sent together as a single query to the DB, returning only filtered results. If the sproc returns results that cannot be mapped to existing entity, then we first create the Entity/EntitySet in the CSDL using Designer, then create a dummy Entity/EntitySet using XML in the SSDL. When creating a EntitySet in the SSDL for this dummy entity, use a TSQL that does not return any results, but does return the relevant columns e.g. select ContactID, FirstName, LastName from dbo.Contact where 1=2 Also insure that the Entity created in the SSDL uses the SQL DataTypes and not .NET DataTypes. If you are unable to open the EDMX file in the designer then note the Errors ... they will give precise info on what is wrong. The Thrid option is to simply create a Native Query in the SSDL using <Function Name="PaymentsforContact" IsComposable="false">   <CommandText>SELECT ActivityId, Activity AS ActivityName, ImagePath, Category FROM dbo.Activities </CommandText></FuncTion> Then map this Function to a existing Entity. This is a quick way to get a custom Entity which is regular Entity with renamed columns or additional columns (which are computed columns). The disadvantage to using this is that It will return all the rows from the Defining query and any filter (if defined) will be applied only at the Client side (after getting all the rows from DB). If you you DefiningQuery instead then we can use that as a Composable Query. The Fourth option (for mapping a READ stored proc results to a non-existent Entity) is to create a View in the Database which returns all the fields that the sproc also returns, then update the Model so that the model contains this View as a Entity. Then map the Read Sproc to this View Entity. The other option would be to simply create the View and remove the sproc altogether. ================================================================ To Execute a SProc that does not return a entity, use a EntityCommand to execute that proc. You cannot call a sproc FunctionImport that does not return Entities From Code, the only way is to use SSDL function calls using EntityCommand.  This changes with EntityFramework Version 4 where you can return Scalar Types, Complex Types, Entities or NonQuery ================================================================ UDF when created as a Function in SSDL, we need to set the Name & IsComposable properties for the Function element. IsComposable is always false for Sprocs, for UDF's set this to true. You cannot call UDF "Function" from within code since you cannot import a UDF Function into the CSDL Model (with Version 1 of EF). only stored procedures can be imported and then mapped to a entity ================================================================ Entity Framework requires properties that are involved in association mappings to be mapped in all of the function mappings for the entity (Insert, Update and Delete). Because Payment has an association to Reservation... hence we need to pass both the paymentId and reservationId to the Delete sproc even though just the paymentId is the PK on the Payment Table. ================================================================ When mapping insert, update and delete procs to a Entity, insure that all the three or none are mapped. Further if you have a base class and derived class in the CSDL, then you must map (ins, upd, del) sprocs to all parent and child entities in the inheritance relationship. Note that this limitation that base and derived entity methods must all must be mapped does not apply when you are mapping Read Stored Procedures.... ================================================================ You can write stored procedures SQL directly into the SSDL by creating a Function element in the SSDL and then once created, you can map this Function to a CSDL Entity directly in the designer during Function Import ================================================================ You can do Entity Splitting such that One Entity maps to multiple tables in the DB. For e.g. the Customer Entity currently derives from Contact Entity...in addition it also references the ContactPersonalInfo Entity. One can copy all properties from the ContactPersonalInfo Entity into the Customer Entity and then Delete the CustomerPersonalInfo entity, finall one needs to map the copied properties to the ContactPersonalInfo Table in Table Mapping (by adding another table (ContactPersonalInfo) to the Table Mapping... this is called Entity Splitting. Thus now when you insert a Customer record, it will automatically create SQL to insert records into the Contact, Customers and ContactPersonalInfo tables even though you have a Single Entity called Customer in the CSDL =================================================================== There is Table by Type Inheritance where another EDM Entity can derive from another EDM entity and absorb the inherted entities properties, for example in the Break Away Geek Adventures EDM, the Customer entity derives (inherits) from the Contact Entity and absorbs all the properties of Contact entity. Thus when you create a Customer Entity in Code and then call context.SaveChanges the Object Context will first create the TSQL to insert into the Contact Table followed by a TSQL to insert into the Customer table =================================================================== Then there is the Table per Hierarchy Inheritance..... where different types are created based on a condition (similar applying a condition to filter a Entity to contain filtered records)... the diference being that the filter condition populates a new Entity Type (derived from the base Entity). In the BreakAway sample the example is Lodging Entity which is a Abstract Entity and Then Resort and NonResort Entities which derive from Lodging Entity and records are filtered based on the value of the Resort Boolean field =================================================================== Then there is Table per Concrete Type Hierarchy where we create a concrete Entity for each table in the database. In the BreakAway sample there is a entity for the Reservation table and another Entity for the OldReservation table even though both the table contain the same number of fields. The OldReservation Entity can then inherit from the Reservation Entity and configure the OldReservation Entity to remove all Scalar Properties from the Entity (since it inherits the properties from Reservation and filters based on ReservationDate field) =================================================================== Complex Types (Complex Properties) Entities in EF can also contain Complex Properties (in addition to Scalar Properties) and these Complex Properties reference a ComplexType (not a EntityType) DropdownList, ListBox, RadioButtonList, CheckboxList, Bulletedlist are examples of List server controls (not data bound controls) these controls cannot use Complex properties during databinding, they need Scalar Properties. So if a Entity contains Complex properties and you need to bind those to list server controls then use projections to return Scalar properties and bind them to the control (the disadvantage is that projected collections are not tracked by the Object Context and hence cannot persist changes to the projected collections bound to controls) ObjectDataSource and EntityDataSource do account for Complex properties and one can bind entities with Complex Properties to Data Source controls and they will be tracked for changes... with no additional plumbing needed to persist changes to these collections bound to controls So DataBound controls like GridView, FormView need to use EntityDataSource or ObjectDataSource as a datasource for entities that contain Complex properties so that changes to the datasource done using the GridView can be persisted to the DB (enabling the controls for updates)....if you cannot use the EntityDataSource you need to flatten the ComplexType Properties using projections With EF Version 4 ComplexTypes are supported by the Designer and can add/remove/compose Complex Types directly using the Designer =================================================================== Conditional Mapping ... is like Table per Hierarchy Inheritance where Entities inherit from a base class and then used conditions to populate the EntitySet (called conditional Mapping). Conditional Mapping has limitations since you can only use =, is null and IS NOT NULL Conditions to do conditional mapping. If you need more operators for filtering/mapping conditionally then use QueryView(or possibly Defining Query) to create a readonly entity. QueryView are readonly by default... the EntitySet created by the QueryView is enabled for change tracking by the ObjectContext, however the ObjectContext cannot create insert/update/delete TSQL statements for these Entities when SaveChanges is called since it is QueryView. One way to get around this limitation is to map stored procedures for the insert/update/delete operations in the Designer. =================================================================== Difference between QueryView and Defining Query : QueryView is defined in the (MSL) Mapping File/section of the EDM XML, whereas the DefiningQuery is defined in the store schema (SSDL). QueryView is written using Entity SQL and is this database agnostic and can be used against any database/Data Layer. DefiningQuery is written using Database Lanaguage i.e. TSQL or PSQL thus you have more control =================================================================== Performance: Lazy loading is deferred loading done automatically. lazy loading is supported with EF version4 and is on by default. If you need to turn it off then use context.ContextOptions.lazyLoadingEnabled = false To improve Performance consider PreCompiling the ObjectQuery using the CompiledQuery.Compile method

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  • Mr Flibble: As Seen Through a Lens, Darkly

    - by Phil Factor
    One of the rewarding things about getting involved with Simple-Talk has been in meeting and working with some pretty daunting talents. I’d like to say that Dom Reed’s talents are at the end of the visible spectrum, but then there is Richard, who pops up on national radio occasionally, presenting intellectual programs, Andrew, master of the ukulele, with his pioneering local history work, and Tony with marathon running and his past as a university lecturer. However, Dom, who is Red Gate’s head of creative design and who did the preliminary design work for Simple-Talk, has taken the art photography to an extreme that was impossible before Photoshop. He’s not the first person to take a photograph of himself every day for two years, but he is definitely the first to weave the results into a frightening narrative that veers from comedy to pathos, using all the arts of Photoshop to create a fictional character, Mr Flibble.   Have a look at some of the Flickr pages. Uncle Spike The B-Men – Woolverine The 2011 BoyZ iN Sink reunion tour turned out to be their last Error 404 – Flibble not found Mr Flibble is not a normal type of alter-ego. We generally prefer to choose bronze age warriors of impossibly magnificent physique and stamina; superheroes who bestride the world, scorning the forces of evil and anarchy in a series noble and righteous quests. Not so Dom, whose Mr Flibble is vulnerable, and laid low by an addiction to toxic substances. His work has gained an international cult following and is used as course material by several courses in photography. Although his work was for a while ignored by the more conventional world of ‘art’ photography they became famous through the internet. His photos have received well over a million views on Flickr. It was definitely time to turn this work into a book, because the whole sequence of images has its maximum effect when seen in sequence. He has a Kickstarter project page, one of the first following the recent UK launch of the crowdfunding platform. The publication of the book should be a major event and the £45 I shall divvy up will be one of the securest investments I shall ever make. The local news in Cambridge picked up on the project and I can quote from the report by the excellent Cabume website , the source of Tech news from the ‘Cambridge cluster’ Put really simply Mr Flibble likes to dress up and take pictures of himself. One of the benefits of a split personality, however is that Mr Flibble is supported in his endeavour by Reed’s top notch photography skills, supreme mastery of Photoshop and unflinching dedication to the cause. The duo have collaborated to take a picture every day for the past 730-plus days. It is not a big surprise that neither Mr Flibble nor Reed watches any TV: In addition to his full-time role at Cambridge software house,Red Gate Software as head of creativity and the two to five hours a day he spends taking the Mr Flibble shots, Reed also helps organise the . And now Reed is using Kickstarter to see if the world is ready for a Mr Flibble coffee table book. Judging by the early response it is. At the time of writing, just a few days after it went live, ‘I Drink Lead Paint: An absurd photography book by Mr Flibble’ had raised £1,545 of the £10,000 target it needs to raise by the Friday 30 November deadline from 37 backers. Following the standard Kickstarter template, Reed is offering a series of rewards based on the amount pledged, ranging from a Mr Flibble desktop wallpaper for pledges of £5 or more to a signed copy of the book for pledges of £45 or more, right up to a starring role in the book for £1,500. Mr Flibble is unquestionably one of the more deranged Kickstarter hopefuls, but don’t think for a second that he doesn’t have a firm grasp on the challenges he faces on the road to immortalisation on 150 gsm stock. Under the section ‘risks and challenges’ on his Kickstarter page his statement begins: “An angry horde of telepathic iguanas discover the world’s last remaining stock of vintage lead paint and hold me to ransom. Gosh how I love to guzzle lead paint. Anyway… faced with such brazen bravado, I cower at the thought of taking on their combined might and die a sad and lonely Flibble deprived of my one and only true liquid love.” At which point, Reed manages to wrestle away the keyboard, giving him the opportunity to present slightly more cogent analysis of the obstacles the project must still overcome. We asked Reed a few questions about Mr Flibble’s Kickstarter adventure and felt that his responses were worth publishing in full: Firstly, how did you manage it – holding down a full time job and also conceiving and executing these ideas on a daily basis? I employed a small team of ferocious gerbils to feed me ideas on a daily basis. Whilst most of their ideas were incomprehensibly rubbish and usually revolved around food, just occasionally they’d give me an idea like my B-Men series. As a backup plan though, I found that the best way to generate ideas was to actually start taking photos. If I were to stand in front of the camera, pull a silly face, place a vegetable on my head or something else equally stupid, the resulting photo of that would typically spark an idea when I came to look at it. Sitting around idly trying to think of an idea was doomed to result in no ideas. I admit that I really struggled with time. I’m proud that I never missed a day, but it was definitely hard when you were late from work, tired or doing something socially on the same day. I don’t watch TV, which I guess really helps, because I’d frequently be spending 2-5 hours taking and processing the photos every day. Are there any overlaps between software development and creative thinking? Software is an inherently creative business and the speed that it moves ensures you always have to find solutions to new things. Everyone in the team needs to be a problem solver. Has it helped me specifically with my photography? Probably. Working within teams that continually need to figure out new stuff keeps the brain feisty I suppose, and I guess I’m continually exposed to a lot of possible sources of inspiration. How specifically will this Kickstarter project allow you to test the commercial appeal of your work and do you plan to get the book into shops? It’s taken a while to be confident saying it, but I know that people like the work that I do. I’ve had well over a million views of my pictures, many humbling comments and I know I’ve garnered some loyal fans out there who anticipate my next photo. For me, this Kickstarter is about seeing if there’s worth to my work beyond just making people smile. In an online world where there’s an abundance of freely available content, can you hope to receive anything from what you do, or would people just move onto the next piece of content if you happen to ask for some support? A book has been the single-most requested thing that people have asked me to produce and it’s something that I feel would showcase my work well. It’s just hard to convince people in the publishing industry just now to take any kind of risk – they’ve been hit hard. If I can show that people would like my work enough to buy a book, then it sends a pretty clear picture that publishers might hear, or it gives me the confidence enough to invest in myself a bit more – hard to do when you’re riddled with self-doubt! I’d love to see my work in the shops, yes. I could see it being the thing that someone flips through idly as they’re Christmas shopping and recognizing that it’d be just the perfect gift for their difficult to buy for friend or relative. That said, working in the software industry means I’m clearly aware of how I could use technology to distribute my work, but I can’t deny that there’s something very appealing to having a physical thing to hold in your hands. If the project is successful is there a chance that it could become a full-time job? At the moment that seems like a distant dream, as should this be successful, there are many more steps I’d need to take to reach any kind of business viability. Kickstarter seems exactly that – a way for people to help kick start me into something that could take off. If people like my work and want me to succeed with it, then taking a look at my Kickstarter page (and hopefully pledging a bit of support) would make my elbows blush considerably. So there is is. An opportunity to open the wallet just a bit to ensure that one of the more unusual talents sees the light in the format it deserves.  

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  • Connecting Linux to WatchGuard Firebox SSL (OpenVPN client)

    Recently, I got a new project assignment that requires to connect permanently to the customer's network through VPN. They are using a so-called SSL VPN. As I am using OpenVPN since more than 5 years within my company's network I was quite curious about their solution and how it would actually be different from OpenVPN. Well, short version: It is a disguised version of OpenVPN. Unfortunately, the company only offers a client for Windows and Mac OS which shouldn't bother any Linux user after all. OpenVPN is part of every recent distribution and can be activated in a couple of minutes - both client as well as server (if necessary). WatchGuard Firebox SSL - About dialog Borrowing some files from a Windows client installation Initially, I didn't know about the product, so therefore I went through the installation on Windows 8. No obstacles (and no restart despite installation of TAP device drivers!) here and the secured VPN channel was up and running in less than 2 minutes or so. Much appreciated from both parties - customer and me. Of course, this whole client package and my long year approved and stable installation ignited my interest to have a closer look at the WatchGuard client. Compared to the original OpenVPN client (okay, I have to admit this is years ago) this commercial product is smarter in terms of file locations during installation. You'll be able to access the configuration and key files below your roaming application data folder. To get there, simply enter '%AppData%\WatchGuard\Mobile VPN' in your Windows/File Explorer and confirm with Enter/Return. This will display the following files: Application folder below user profile with configuration and certificate files From there we are going to borrow four files, namely: ca.crt client.crt client.ovpn client.pem and transfer them to the Linux system. You might also be able to isolate those four files from a Mac OS client. Frankly, I'm just too lazy to run the WatchGuard client installation on a Mac mini only to find the folder location, and I'm going to describe why a little bit further down this article. I know that you can do that! Feedback in the comment section is appreciated. Configuration of OpenVPN (console) Depending on your distribution the following steps might be a little different but in general you should be able to get the important information from it. I'm going to describe the steps in Ubuntu 13.04 (Raring Ringtail). As usual, there are two possibilities to achieve your goal: console and UI. Let's what it is necessary to be done. First of all, you should ensure that you have OpenVPN installed on your system. Open your favourite terminal application and run the following statement: $ sudo apt-get install openvpn network-manager-openvpn network-manager-openvpn-gnome Just to be on the safe side. The four above mentioned files from your Windows machine could be copied anywhere but either you place them below your own user directory or you put them (as root) below the default directory: /etc/openvpn At this stage you would be able to do a test run already. Just in case, run the following command and check the output (it's the similar information you would get from the 'View Logs...' context menu entry in Windows: $ sudo openvpn --config client.ovpn Pay attention to the correct path to your configuration and certificate files. OpenVPN will ask you to enter your Auth Username and Auth Password in order to establish the VPN connection, same as the Windows client. Remote server and user authentication to establish the VPN Please complete the test run and see whether all went well. You can disconnect pressing Ctrl+C. Simplifying your life - authentication file In my case, I actually set up the OpenVPN client on my gateway/router. This establishes a VPN channel between my network and my client's network and allows me to switch machines easily without having the necessity to install the WatchGuard client on each and every machine. That's also very handy for my various virtualised Windows machines. Anyway, as the client configuration, key and certificate files are located on a headless system somewhere under the roof, it is mandatory to have an automatic connection to the remote site. For that you should first change the file extension '.ovpn' to '.conf' which is the default extension on Linux systems for OpenVPN, and then open the client configuration file in order to extend an existing line. $ sudo mv client.ovpn client.conf $ sudo nano client.conf You should have a similar content to this one here: dev tunclientproto tcp-clientca ca.crtcert client.crtkey client.pemtls-remote "/O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server"remote-cert-eku "TLS Web Server Authentication"remote 1.2.3.4 443persist-keypersist-tunverb 3mute 20keepalive 10 60cipher AES-256-CBCauth SHA1float 1reneg-sec 3660nobindmute-replay-warningsauth-user-pass auth.txt Note: I changed the IP address of the remote directive above (which should be obvious, right?). Anyway, the required change is marked in red and we have to create a new authentication file 'auth.txt'. You can give the directive 'auth-user-pass' any file name you'd like to. Due to my existing OpenVPN infrastructure my setup differs completely from the above written content but for sake of simplicity I just keep it 'as-is'. Okay, let's create this file 'auth.txt' $ sudo nano auth.txt and just put two lines of information in it - username on the first, and password on the second line, like so: myvpnusernameverysecretpassword Store the file, change permissions, and call openvpn with your configuration file again: $ sudo chmod 0600 auth.txt $ sudo openvpn --config client.conf This should now work without being prompted to enter username and password. In case that you placed your files below the system-wide location /etc/openvpn you can operate your VPNs also via service command like so: $ sudo service openvpn start client $ sudo service openvpn stop client Using Network Manager For newer Linux users or the ones with 'console-phobia' I'm going to describe now how to use Network Manager to setup the OpenVPN client. For this move your mouse to the systray area and click on Network Connections => VPN Connections => Configure VPNs... which opens your Network Connections dialog. Alternatively, use the HUD and enter 'Network Connections'. Network connections overview in Ubuntu Click on 'Add' button. On the next dialog select 'Import a saved VPN configuration...' from the dropdown list and click on 'Create...' Choose connection type to import VPN configuration Now you navigate to your folder where you put the client files from the Windows system and you open the 'client.ovpn' file. Next, on the tab 'VPN' proceed with the following steps (directives from the configuration file are referred): General Check the IP address of Gateway ('remote' - we used 1.2.3.4 in this setup) Authentication Change Type to 'Password with Certificates (TLS)' ('auth-pass-user') Enter User name to access your client keys (Auth Name: myvpnusername) Enter Password (Auth Password: verysecretpassword) and choose your password handling Browse for your User Certificate ('cert' - should be pre-selected with client.crt) Browse for your CA Certificate ('ca' - should be filled as ca.crt) Specify your Private Key ('key' - here: client.pem) Then click on the 'Advanced...' button and check the following values: Use custom gateway port: 443 (second value of 'remote' directive) Check the selected value of Cipher ('cipher') Check HMAC Authentication ('auth') Enter the Subject Match: /O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server ('tls-remote') Finally, you have to confirm and close all dialogs. You should be able to establish your OpenVPN-WatchGuard connection via Network Manager. For that, click on the 'VPN Connections => client' entry on your Network Manager in the systray. It is advised that you keep an eye on the syslog to see whether there are any problematic issues that would require some additional attention. Advanced topic: routing As stated above, I'm running the 'WatchGuard client for Linux' on my head-less server, and since then I'm actually establishing a secure communication channel between two networks. In order to enable your network clients to get access to machines on the remote side there are two possibilities to enable that: Proper routing on both sides of the connection which enables both-direction access, or Network masquerading on the 'client side' of the connection Following, I'm going to describe the second option a little bit more in detail. The Linux system that I'm using is already configured as a gateway to the internet. I won't explain the necessary steps to do that, and will only focus on the additional tweaks I had to do. You can find tons of very good instructions and tutorials on 'How to setup a Linux gateway/router' - just use Google. OK, back to the actual modifications. First, we need to have some information about the network topology and IP address range used on the 'other' side. We can get this very easily from /var/log/syslog after we established the OpenVPN channel, like so: $ sudo tail -n20 /var/log/syslog Or if your system is quite busy with logging, like so: $ sudo less /var/log/syslog | grep ovpn The output should contain PUSH received message similar to the following one: Jul 23 23:13:28 ios1 ovpn-client[789]: PUSH: Received control message: 'PUSH_REPLY,topology subnet,route 192.168.1.0 255.255.255.0,dhcp-option DOMAIN ,route-gateway 192.168.6.1,topology subnet,ping 10,ping-restart 60,ifconfig 192.168.6.2 255.255.255.0' The interesting part for us is the route command which I highlighted already in the sample PUSH_REPLY. Depending on your remote server there might be multiple networks defined (172.16.x.x and/or 10.x.x.x). Important: The IP address range on both sides of the connection has to be different, otherwise you will have to shuffle IPs or increase your the netmask. {loadposition content_adsense} After the VPN connection is established, we have to extend the rules for iptables in order to route and masquerade IP packets properly. I created a shell script to take care of those steps: #!/bin/sh -eIPTABLES=/sbin/iptablesDEV_LAN=eth0DEV_VPNS=tun+VPN=192.168.1.0/24 $IPTABLES -A FORWARD -i $DEV_LAN -o $DEV_VPNS -d $VPN -j ACCEPT$IPTABLES -A FORWARD -i $DEV_VPNS -o $DEV_LAN -s $VPN -j ACCEPT$IPTABLES -t nat -A POSTROUTING -o $DEV_VPNS -d $VPN -j MASQUERADE I'm using the wildcard interface 'tun+' because I have multiple client configurations for OpenVPN on my server. In your case, it might be sufficient to specify device 'tun0' only. Simplifying your life - automatic connect on boot Now, that the client connection works flawless, configuration of routing and iptables is okay, we might consider to add another 'laziness' factor into our setup. Due to kernel updates or other circumstances it might be necessary to reboot your system. Wouldn't it be nice that the VPN connections are established during the boot procedure? Yes, of course it would be. To achieve this, we have to configure OpenVPN to automatically start our VPNs via init script. Let's have a look at the responsible 'default' file and adjust the settings accordingly. $ sudo nano /etc/default/openvpn Which should have a similar content to this: # This is the configuration file for /etc/init.d/openvpn## Start only these VPNs automatically via init script.# Allowed values are "all", "none" or space separated list of# names of the VPNs. If empty, "all" is assumed.# The VPN name refers to the VPN configutation file name.# i.e. "home" would be /etc/openvpn/home.conf#AUTOSTART="all"#AUTOSTART="none"#AUTOSTART="home office"## ... more information which remains unmodified ... With the OpenVPN client configuration as described above you would either set AUTOSTART to "all" or to "client" to enable automatic start of your VPN(s) during boot. You should also take care that your iptables commands are executed after the link has been established, too. You can easily test this configuration without reboot, like so: $ sudo service openvpn restart Enjoy stable VPN connections between your Linux system(s) and a WatchGuard Firebox SSL remote server. Cheers, JoKi

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