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  • Hosting and consuming WCF services without configuration files

    - by martinsj
    In this post, I'll demonstrate how to configure both the host and the client in code without the need for configuring services i the <system.serviceModel> section of the config-file. In fact, you don't need a  <system.serviceModel> section at all. What you'll do need (and want) sometimes, is the Uri of the service in the configuration file. Configuring the Uri of the the service is actually only needed for the client or when self-hosting, not when hosting in IIS. So, exactly What do we need to configure? The binding type and the binding constraints The metadata behavior Debug behavior You can of course configure even more, and even more if you want to, WCF is after all the king of configuration… As an example I'll be hosting and consuming a service that removes most of the default constraints for WCF-services, using a BasicHttpBinding. Of course, in regards to security, it is probably better to have some constraints on the server, but this is only a demonstration. The ServerConfig class in the code beneath is a static helper class that will be used in the examples. In this post, I’ll be using this helper-class for all configuration, for both the server and the client. In WCF, the  client and the server have both their own WCF-configuration. With this piece of code, they will be sharing the same configuration. 1: public static class ServiceConfig 2: { 3: public static Binding DefaultBinding 4: { 5: get 6: { 7: var binding = new BasicHttpBinding(); 8: Configure(binding); 9: return binding; 10: } 11: } 12:  13: public static void Configure(HttpBindingBase binding) 14: { 15: if (binding == null) 16: { 17: throw new ArgumentException("Argument 'binding' cannot be null. Cannot configure binding."); 18: } 19:  20: binding.SendTimeout = new TimeSpan(0, 0, 30, 0); // 30 minute timeout 21: binding.MaxBufferSize = Int32.MaxValue; 22: binding.MaxBufferPoolSize = 2147483647; 23: binding.MaxReceivedMessageSize = Int32.MaxValue; 24: binding.ReaderQuotas.MaxArrayLength = Int32.MaxValue; 25: binding.ReaderQuotas.MaxBytesPerRead = Int32.MaxValue; 26: binding.ReaderQuotas.MaxDepth = Int32.MaxValue; 27: binding.ReaderQuotas.MaxNameTableCharCount = Int32.MaxValue; 28: binding.ReaderQuotas.MaxStringContentLength = Int32.MaxValue; 29: } 30:  31: public static ServiceMetadataBehavior ServiceMetadataBehavior 32: { 33: get 34: { 35: return new ServiceMetadataBehavior 36: { 37: HttpGetEnabled = true, 38: MetadataExporter = {PolicyVersion = PolicyVersion.Policy15} 39: }; 40: } 41: } 42:  43: public static ServiceDebugBehavior ServiceDebugBehavior 44: { 45: get 46: { 47: var smb = new ServiceDebugBehavior(); 48: Configure(smb); 49: return smb; 50: } 51: } 52:  53:  54: public static void Configure(ServiceDebugBehavior behavior) 55: { 56: if (behavior == null) 57: { 58: throw new ArgumentException("Argument 'behavior' cannot be null. Cannot configure debug behavior."); 59: } 60: 61: behavior.IncludeExceptionDetailInFaults = true; 62: } 63: } Configuring the server There are basically two ways to host a WCF service, in IIS and self-hosting. When hosting a WCF service in a production environment using SOA architecture, you'll be most likely hosting it in IIS. When testing the service in integration tests, it's very handy to be able to self-host services in the unit-tests. In fact, you can share the the WCF configuration for self-hosted services and services hosted in IIS. And that is exactly what you want to do, testing the same configurations for test and production environments.   Configuring when Self-hosting When self-hosting, in order to start the service, you'll have to instantiate the ServiceHost class, configure the  service and open it. 1: // Create the service-host. 2: var host = new ServiceHost(typeof(MyService), endpoint); 3:  4: // Configure the binding 5: host.AddServiceEndpoint(typeof(IMyService), ServiceConfig.DefaultBinding, endpoint); 6:  7: // Configure metadata behavior 8: host.Description.Behaviors.Add(ServiceConfig.ServiceMetadataBehavior); 9:  10: // Configure debgug behavior 11: ServiceConfig.Configure((ServiceDebugBehavior)host.Description.Behaviors[typeof(ServiceDebugBehavior)]); 12: 13: // Start listening to the service 14: host.Open(); 15:  Configuring when hosting in IIS When you create a WCF service application with the wizard in Visual Studio, you'll end up with bits and pieces of code in order to get the service running: Svc-file with codebehind. A interface to the service Web.config In order to get rid of the configuration in the <system.serviceModel> section, which the wizard has generated for us, we must tell the service that we have a factory that will create the service for us. We do this by changing the markup for the svc-file: 1: <%@ ServiceHost Language="C#" Debug="true" Service="Namespace.MyService" Factory="Namespace.ServiceHostFactory" %> The markup tells IIS that we have a factory called ServiceHostFactory for this service. The service factory has a method we can override which will be called when someone asks IIS for the service. There are overloads we can override: 1: System.ServiceModel.ServiceHostBase CreateServiceHost(string constructorString, Uri[] baseAddresses) 2: System.ServiceModel.ServiceHost CreateServiceHost(Type serviceType, Uri[] baseAddresses) 3:  In this example, we'll be using the last one, so our implementation looks like this: 1: public class ServiceHostFactory : System.ServiceModel.Activation.ServiceHostFactory 2: { 3:  4: protected override System.ServiceModel.ServiceHost CreateServiceHost(Type serviceType, Uri[] baseAddresses) 5: { 6: var host = base.CreateServiceHost(serviceType, baseAddresses); 7: host.Description.Behaviors.Add(ServiceConfig.ServiceMetadataBehavior); 8: ServiceConfig.Configure((ServiceDebugBehavior)host.Description.Behaviors[typeof(ServiceDebugBehavior)]); 9: return host; 10: } 11: } 12:  1: public class ServiceHostFactory : System.ServiceModel.Activation.ServiceHostFactory 2: { 3: 4: protected override System.ServiceModel.ServiceHost CreateServiceHost(Type serviceType, Uri[] baseAddresses) 5: { 6: var host = base.CreateServiceHost(serviceType, baseAddresses); 7: host.Description.Behaviors.Add(ServiceConfig.ServiceMetadataBehavior); 8: ServiceConfig.Configure((ServiceDebugBehavior)host.Description.Behaviors[typeof(ServiceDebugBehavior)]); 9: return host; 10: } 11: } 12: As you can see, we are using the same configuration helper we used when self-hosting. Now, when you have a factory, the <system.serviceModel> section of the configuration can be removed, because the section will be ignored when the service has a custom factory. If you want to configure something else in the config-file, one could configure in some other section.   Configuring the client Microsoft has helpfully created a ChannelFactory class in order to create a proxy client. When using this approach, you don't have generate those awfull proxy classes for the client. If you share the contracts with the server in it's own assembly like in the layer diagram under, you can share the same piece of code. The contracts in WCF are the interface to the service and if any, the datacontracts (custom types) the service depends on. Using the ChannelFactory with our configuration helper-class is very simple: 1: var identity = EndpointIdentity.CreateDnsIdentity("localhost"); 2: var endpointAddress = new EndpointAddress(endPoint, identity); 3: var factory = new ChannelFactory<IMyService>(DeployServiceConfig.DefaultBinding, endpointAddress); 4: using (var myService = new factory.CreateChannel()) 5: { 6: myService.Hello(); 7: } 8: factory.Close();   Happy configuration!

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  • “It’s only test code…”

    - by Chris George
    “Let me hack this in, it’s only test code”, “Don’t worry about getting it reviewed, it’s only test code”, “It doesn’t have to be elegant or efficient, it’s only test code”… do these phrases sound familiar? Chances are if you’ve working with test automation, at one point or other you will have heard these phrases, you have probably even used them yourself! What is certain is that code written under this “it’s only test code” mantra will come back and bite you in the arse! I’ve recently encountered a case where a test was giving a false positive, therefore hiding a real product bug because that test code was very badly written. Firstly it was very difficult to understand what the test was actually trying to achieve let alone how it was doing it, and this complexity masked a simple logic error. These issues are real and they do happen. Let’s take a step back from this and look at what we are trying to do. We are writing test code that tests product code, and we do this to create a suite of tests that will help protect our software against regressions. This test code is making sure that the product behaves as it should by employing some sort of expected result verification. The simple cases of these are generally not a problem. However, automation allows us to explore more complex scenarios in many more permutations. As this complexity increases then so does the complexity of the test code. It is at this point that code which has not been architected properly will cause problems.   Keep your friends close… So, how do we make sure we are doing it right? The development teams I have worked on have always had Test Engineers working very closely with their Software Engineers. This is something that I have always tried to take full advantage of. They are coding experts! So run your ideas past them, ask for advice on how to structure your code, help you design your data structures. This may require a shift in your teams viewpoint, as contrary to this section title and folklore, Software Engineers are not actually the mortal enemy of Test Engineers. As time progresses, and test automation becomes more and more ingrained in what we do, the two roles are converging more than ever. Over the 16 years I have spent as a Test Engineer, I have seen the grey area between the two roles grow significantly larger. This serves to strengthen the relationship and common bond between the two roles which helps to make test code activities so much easier!   Pair for the win Possibly the best thing you could do to write good test code is to pair program on the task. This will serve a few purposes. you will get the benefit of the Software Engineers knowledge and experience the Software Engineer will gain knowledge on the testing process. Sharing the love is a wonderful thing! two pairs of eyes are always better than one… And so are two brains. Between the two of you, I will guarantee you will derive more useful test cases than if it was just one of you.   Code reviews Another policy which certainly pays dividends is the practice of code reviews. By having one of your peers review your code before you commit it serves two purposes. Firstly, it forces you to explain your code. Just the act of doing this will often pick up errors in your code. Secondly, it gets yet another pair of eyes on your code! I cannot stress enough how important code reviews are. The benefits they offer apply as much to product code as test code. In short, Software and Test Engineers should all be doing them! It can be extended even further by getting test code reviewed by a Software Engineer and a Test Engineer, and likewise product code. This serves to keep both functions in the loop with changes going on within your code base.   Learn from your devs I briefly touched on this earlier but I’d like to go into more detail here. Pairing with your Software Engineers when writing your test code is such an amazing opportunity to improve your coding skills. As I sit here writing this article waiting to be called into court for jury service, it reminds me that it takes a lot of patience to be a Test Engineer, almost as much as it takes to be a juror! However tempting it is to go rushing in and start writing your automated tests, resist that urge. Discuss what you want to achieve then talk through the approach you’re going to take. Then code it up together. I find it really enlightening to ask questions like ‘is there a better way to do this?’ Or ‘is this how you would code it?’ The latter question, especially, is where I learn the most. I’ve found that most Software Engineers will be reluctant to show you the ‘right way’ to code something when writing tests because they perceive the ‘right way’ to be too complicated for the Test Engineer (e.g. not mentioning LINQ and instead doing something verbose). So by asking how THEY would code it, it unleashes their true dev-ness and advanced code usually ensues! I would like to point out, however, that you don’t have to accept their method as the final answer. On numerous occasions I have opted for the more simple/verbose solution because I found the code written by the Software Engineer too advanced and therefore I would find it unreadable when I return to the code in a months’ time! Always keep the target audience in mind when writing clever code, and in my case that is mostly Test Engineers.  

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  • REST to Objects in C#

    RESTful interfaces for web services are all the rage for many Web 2.0 sites.  If you want to consume these in a very simple fashion, LINQ to XML can do the job pretty easily in C#.  If you go searching for help on this, youll find a lot of incomplete solutions and fairly large toolkits and frameworks (guess how I know this) this quick article is meant to be a no fluff just stuff approach to making this work. POCO Objects Lets assume you have a Model that you want to suck data into from a RESTful web service.  Ideally this is a Plain Old CLR Object, meaning it isnt infected with any persistence or serialization goop.  It might look something like this: public class Entry { public int Id; public int UserId; public DateTime Date; public float Hours; public string Notes; public bool Billable;   public override string ToString() { return String.Format("[{0}] User: {1} Date: {2} Hours: {3} Notes: {4} Billable {5}", Id, UserId, Date, Hours, Notes, Billable); } } Not that this isnt a completely trivial object.  Lets look at the API for the service.  RESTful HTTP Service In this case, its TickSpots API, with the following sample output: <?xml version="1.0" encoding="UTF-8"?> <entries type="array"> <entry> <id type="integer">24</id> <task_id type="integer">14</task_id> <user_id type="integer">3</user_id> <date type="date">2008-03-08</date> <hours type="float">1.00</hours> <notes>Had trouble with tribbles.</notes> <billable>true</billable> # Billable is an attribute inherited from the task <billed>true</billed> # Billed is an attribute to track whether the entry has been invoiced <created_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</created_at> <updated_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</updated_at> # The following attributes are derived and provided for informational purposes: <user_email>[email protected]</user_email> <task_name>Remove converter assembly</task_name> <sum_hours type="float">2.00</sum_hours> <budget type="float">10.00</budget> <project_name>Realign dilithium crystals</project_name> <client_name>Starfleet Command</client_name> </entry> </entries> Im assuming in this case that I dont necessarily care about all of the data fields the service is returning I just need some of them for my applications purposes.  Thus, you can see there are more elements in the <entry> XML than I have in my Entry class. Get The XML with C# The next step is to get the XML.  The following snippet does the heavy lifting once you pass it the appropriate URL: protected XElement GetResponse(string uri) { var request = WebRequest.Create(uri) as HttpWebRequest; request.UserAgent = ".NET Sample"; request.KeepAlive = false;   request.Timeout = 15 * 1000;   var response = request.GetResponse() as HttpWebResponse;   if (request.HaveResponse == true && response != null) { var reader = new StreamReader(response.GetResponseStream()); return XElement.Parse(reader.ReadToEnd()); } throw new Exception("Error fetching data."); } This is adapted from the Yahoo Developer article on Web Service REST calls.  Once you have the XML, the last step is to get the data back as your POCO. Use LINQ-To-XML to Deserialize POCOs from XML This is done via the following code: public IEnumerable<Entry> List(DateTime startDate, DateTime endDate) { string additionalParameters = String.Format("start_date={0}&end_date={1}", startDate.ToShortDateString(), endDate.ToShortDateString()); string uri = BuildUrl("entries", additionalParameters);   XElement elements = GetResponse(uri);   var entries = from e in elements.Elements() where e.Name.LocalName == "entry" select new Entry { Id = int.Parse(e.Element("id").Value), UserId = int.Parse(e.Element("user_id").Value), Date = DateTime.Parse(e.Element("date").Value), Hours = float.Parse(e.Element("hours").Value), Notes = e.Element("notes").Value, Billable = bool.Parse(e.Element("billable").Value) }; return entries; }   For completeness, heres the BuildUrl method for my TickSpot API wrapper: // Change these to your settings protected const string projectDomain = "DOMAIN.tickspot.com"; private const string authParams = "[email protected]&password=MyTickSpotPassword";   protected string BuildUrl(string apiMethod, string additionalParams) { if (projectDomain.Contains("DOMAIN")) { throw new ApplicationException("You must update your domain in ProjectRepository.cs."); } if (authParams.Contains("MyTickSpotPassword")) { throw new ApplicationException("You must update your email and password in ProjectRepository.cs."); } return string.Format("https://{0}/api/{1}?{2}&{3}", projectDomain, apiMethod, authParams, additionalParams); } Thats it!  Now go forth and consume XML and map it to classes you actually want to work with.  Have fun! Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • ANTS CLR and Memory Profiler In Depth Review (Part 2 of 2 &ndash; Memory Profiler)

    - by ToStringTheory
    One of the things that people might not know about me, is my obsession to make my code as efficient as possible. Many people might not realize how much of a task or undertaking that this might be, but it is surely a task as monumental as climbing Mount Everest, except this time it is a challenge for the mind… In trying to make code efficient, there are many different factors that play a part – size of project or solution, tiers, language used, experience and training of the programmer, technologies used, maintainability of the code – the list can go on for quite some time. I spend quite a bit of time when developing trying to determine what is the best way to implement a feature to accomplish the efficiency that I look to achieve. One program that I have recently come to learn about – Red Gate ANTS Performance (CLR) and Memory profiler gives me tools to accomplish that job more efficiently as well. In this review, I am going to cover some of the features of the ANTS memory profiler set by compiling some hideous example code to test against. Notice As a member of the Geeks With Blogs Influencers program, one of the perks is the ability to review products, in exchange for a free license to the program. I have not let this affect my opinions of the product in any way, and Red Gate nor Geeks With Blogs has tried to influence my opinion regarding this product in any way. Introduction – Part 2 In my last post, I reviewed the feature packed Red Gate ANTS Performance Profiler.  Separate from the Red Gate Performance Profiler is the Red Gate ANTS Memory Profiler – a simple, easy to use utility for checking how your application is handling memory management…  A tool that I wish I had had many times in the past.  This post will be focusing on the ANTS Memory Profiler and its tool set. The memory profiler has a large assortment of features just like the Performance Profiler, with the new session looking nearly exactly alike: ANTS Memory Profiler Memory profiling is not something that I have to do very often…  In the past, the few cases I’ve had to find a memory leak in an application I have usually just had to trace the code of the operations being performed to look for oddities…  Sadly, I have come across more undisposed/non-using’ed IDisposable objects, usually from ADO.Net than I would like to ever see.  Support is not fun, however using ANTS Memory Profiler makes this task easier.  For this round of testing, I am going to use the same code from my previous example, using the WPF application. This time, I will choose the ‘Profile Memory’ option from the ANTS menu in Visual Studio, which launches the solution in its currently configured state/start-up project, and then launches the ANTS Memory Profiler to help.  It prepopulates all of the fields with the current project information, and all I have to do is select the ‘Start Profiling’ option. When the window comes up, it is actually quite barren, just giving ideas on how to work the profiler.  You start by getting to the point in your application that you want to profile, and then taking a ‘Memory Snapshot’.  This performs a full garbage collection, and snapshots the managed heap.  Using the same WPF app as before, I will go ahead and take a snapshot now. As you can see, ANTS is already giving me lots of information regarding the snapshot, however this is just a snapshot.  The whole point of the profiler is to perform an action, usually one where a memory problem is being noticed, and then take another snapshot and perform a diff between them to see what has changed.  I am going to go ahead and generate 5000 primes, and then take another snapshot: As you can see, ANTS is already giving me a lot of new information about this snapshot compared to the last.  Information such as difference in memory usage, fragmentation, class usage, etc…  If you take more snapshots, you can use the dropdown at the top to set your actual comparison snapshots. If you beneath the timeline, you will see a breadcrumb trail showing how best to approach profiling memory using ANTS.  When you first do the comparison, you start on the Summary screen.  You can either use the charts at the bottom, or switch to the class list screen to get to the next step.  Here is the class list screen: As you can see, it lists information about all of the instances between the snapshots, as well as at the bottom giving you a way to filter by telling ANTS what your problem is.  I am going to go ahead and select the Int16[] to look at the Instance Categorizer Using the instance categorizer, you can travel backwards to see where all of the instances are coming from.  It may be hard to see in this image, but hopefully the lightbox (click on it) will help: I can see that all of these instances are rooted to the application through the UI TextBlock control.  This image will probably be even harder to see, however using the ‘Instance Retention Graph’, you can trace an objects memory inheritance up the chain to see its roots as well.  This is a simple example, as this is simply a known element.  Usually you would be profiling an actual problem, and comparing those differences.  I know in the past, I have spotted a problem where a new context was created per page load, and it was rooted into the application through an event.  As the application began to grow, performance and reliability problems started to emerge.  A tool like this would have been a great way to identify the problem quickly. Overview Overall, I think that the Red Gate ANTS Memory Profiler is a great utility for debugging those pesky leaks.  3 Biggest Pros: Easy to use interface with lots of options for configuring profiling session Intuitive and helpful interface for drilling down from summary, to instance, to root graphs ANTS provides an API for controlling the profiler. Not many options, but still helpful. 2 Biggest Cons: Inability to automatically snapshot the memory by interval Lack of complete integration with Visual Studio via an extension panel Ratings Ease of Use (9/10) – I really do believe that they have brought simplicity to the once difficult task of memory profiling.  I especially liked how it stepped you further into the drilldown by directing you towards the best options. Effectiveness (10/10) – I believe that the profiler does EXACTLY what it purports to do.  Features (7/10) – A really great set of features all around in the application, however, I would like to see some ability for automatically triggering snapshots based on intervals or framework level items such as events. Customer Service (10/10) – My entire experience with Red Gate personnel has been nothing but good.  their people are friendly, helpful, and happy! UI / UX (9/10) – The interface is very easy to get around, and all of the options are easy to find.  With a little bit of poking around, you’ll be optimizing Hello World in no time flat! Overall (9/10) – Overall, I am happy with the Memory Profiler and its features, as well as with the service I received when working with the Red Gate personnel.  Thank you for reading up to here, or skipping ahead – I told you it would be shorter!  Please, if you do try the product, drop me a message and let me know what you think!  I would love to hear any opinions you may have on the product. Code Feel free to download the code I used above – download via DropBox

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  • Making Those PanelBoxes Behave

    - by Duncan Mills
    I have a little problem to solve earlier this week - misbehaving <af:panelBox> components... What do I mean by that? Well here's the scenario, I have a page fragment containing a set of panelBoxes arranged vertically. As it happens, they are stamped out in a loop but that does not really matter. What I want to be able to do is to provide the user with a simple UI to close and open all of the panelBoxes in concert. This could also apply to showDetailHeader and similar items with a disclosed attrubute, but in this case it's good old panelBoxes.  Ok, so the basic solution to this should be self evident. I can set up a suitable scoped managed bean that the panelBoxes all refer to for their disclosed attribute state. Then the open all / close commandButtons in the UI can simply set the state of that bean for all the panelBoxes to pick up via EL on their disclosed attribute. Sound OK? Well that works basically without a hitch, but turns out that there is a slight problem and this is where the framework is attempting to be a little too helpful. The issue is that is the user manually discloses or hides a panelBox then that will override the value that the EL is setting. So for example. I start the page with all panelBoxes collapsed, all set by the EL state I'm storing on the session I manually disclose panelBox no 1. I press the Expand All button - all works as you would hope and all the panelBoxes are now disclosed, including of course panelBox 1 which I just expanded manually. Finally I press the Collapse All button and everything collapses except that first panelBox that I manually disclosed.  The problem is that the component remembers this manual disclosure and that overrides the value provided by the expression. If I change the viewId (navigate away and back) then the panelBox will start to behave again, until of course I touch it again! Now, the more astute amoungst you would think (as I did) Ah, sound like the MDS personalizaton stuff is getting in the way and the solution should simply be to set the dontPersist attribute to disclosed | ALL. Alas this does not fix the issue.  After a little noodling on the best way to approach this I came up with a solution that works well, although if you think of an alternative way do let me know. The principle is simple. In the disclosureListener for the panelBox I take a note of the clientID of the panelBox component that has been touched by the user along with the state. This all gets stored in a Map of Booleans in ViewScope which is keyed by clientID and stores the current disclosed state in the Boolean value.  The listener looks like this (it's held in a request scope backing bean for the page): public void handlePBDisclosureEvent(DisclosureEvent disclosureEvent) { String clientId = disclosureEvent.getComponent().getClientId(FacesContext.getCurrentInstance()); boolean state = disclosureEvent.isExpanded(); pbState.addTouchedPanelBox(clientId, state); } The pbState variable referenced here is a reference to the bean which will hold the state of the panelBoxes that lives in viewScope (recall that everything is re-set when the viewid is changed so keeping this in viewScope is just fine and cleans things up automatically). The addTouchedPanelBox() method looks like this: public void addTouchedPanelBox(String clientId, boolean state) { //create the cache if needed this is just a Map<String,Boolean> if (_touchedPanelBoxState == null) { _touchedPanelBoxState = new HashMap<String, Boolean>(); } // Simply put / replace _touchedPanelBoxState.put(clientId, state); } So that's the first part, we now have a record of every panelBox that the user has touched. So what do we do when the Collapse All or Expand All buttons are pressed? Here we do some JavaScript magic. Basically for each clientID that we have stored away, we issue a client side disclosure event from JavaScript - just as if the user had gone back and changed it manually. So here's the Collapse All button action: public String CloseAllAction() { submitDiscloseOverride(pbState.getTouchedClientIds(true), false); _uiManager.closeAllBoxes(); return null; }  The _uiManager.closeAllBoxes() method is just manipulating the master-state that all of the panelBoxes are bound to using EL. The interesting bit though is the line:  submitDiscloseOverride(pbState.getTouchedClientIds(true), false); To break that down, the first part is a call to that viewScoped state holder to ask for a list of clientIDs that need to be "tweaked": public String getTouchedClientIds(boolean targetState) { StringBuilder sb = new StringBuilder(); if (_touchedPanelBoxState != null && _touchedPanelBoxState.size() > 0) { for (Map.Entry<String, Boolean> entry : _touchedPanelBoxState.entrySet()) { if (entry.getValue() == targetState) { if (sb.length() > 0) { sb.append(','); } sb.append(entry.getKey()); } } } return sb.toString(); } You'll notice that this method only processes those panelBoxes that will be in the wrong state and returns those as a comma separated list. This is then processed by the submitDiscloseOverride() method: private void submitDiscloseOverride(String clientIdList, boolean targetDisclosureState) { if (clientIdList != null && clientIdList.length() > 0) { FacesContext fctx = FacesContext.getCurrentInstance(); StringBuilder script = new StringBuilder(); script.append("overrideDiscloseHandler('"); script.append(clientIdList); script.append("',"); script.append(targetDisclosureState); script.append(");"); Service.getRenderKitService(fctx, ExtendedRenderKitService.class).addScript(fctx, script.toString()); } } This method constructs a JavaScript command to call a routine called overrideDiscloseHandler() in a script attached to the page (using the standard <af:resource> tag). That method parses out the list of clientIDs and sends the correct message to each one: function overrideDiscloseHandler(clientIdList, newState) { AdfLogger.LOGGER.logMessage(AdfLogger.INFO, "Disclosure Hander newState " + newState + " Called with: " + clientIdList); //Parse out the list of clientIds var clientIdArray = clientIdList.split(','); for (var i = 0; i < clientIdArray.length; i++){ var panelBox = flipPanel = AdfPage.PAGE.findComponentByAbsoluteId(clientIdArray[i]); if (panelBox.getComponentType() == "oracle.adf.RichPanelBox"){ panelBox.broadcast(new AdfDisclosureEvent(panelBox, newState)); } }  }  So there you go. You can see how, with a few tweaks the same code could be used for other components with disclosure that might suffer from the same problem, although I'd point out that the behavior I'm working around here us usually desirable. You can download the running example (11.1.2.2) from here. 

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  • SOA Implementation Challenges

    Why do companies think that if they put up a web service that they are doing Service-Oriented Architecture (SOA)? Unfortunately, the IT and business world love to run on the latest hype or buzz words of which very few even understand the meaning. One of the largest issues companies have today as they consider going down the path of SOA, is the lack of knowledge regarding the architectural style and the over usage of the term SOA. So how do we solve this issue?I am sure most of you are thinking by now that you know what SOA is because you developed a few web services.  Isn’t that SOA, right? No, that is not SOA, but instead Just Another Web Service (JAWS). For us to better understand what SOA is let’s look at a few definitions.Douglas K. Bary defines service-oriented architecture as a collection of services. These services are enabled to communicate with each other in order to pass data or coordinating some activity with other services.If you look at this definition closely you will notice that Bary states that services communicate with each other. Let us compare this statement with my first statement regarding companies that claim to be doing SOA when they have just a collection of web services. In order for these web services to for an SOA application they need to be interdependent on one another forming some sort of architectural hierarchy. Just because a company has a few web services does not mean that they are all interconnected.SearchSOA from TechTarget.com states that SOA defines how two computing entities work collectively to enable one entity to perform a unit of work on behalf of another. Once again, just because a company has a few web services does not guarantee that they are even working together let alone if they are performing work for each other.SearchSOA also points out service interactions should be self-contained and loosely-coupled so that all interactions operate independent of each other.Of all the definitions regarding SOA Thomas Erl’s seems to shed the most light on this concept. He states that “SOA establishes an architectural model that aims to enhance the efficiency, agility, and productivity of an enterprise by positioning services as the primary means through which solution logic is represented in support of the realization of the strategic goals associated with service-oriented computing.” (Erl, 2011) Once again this definition proves that a collection of web services does not mean that a company is doing SOA. However, it does mean that a company has a collection of web services, and that is it.In order for a company to start to go down the path of SOA, they must take  a hard look at their existing business process while abstracting away any technology so that they can define what is they really want to accomplish. Once a company has done this, they can begin to factor out common sub business process like credit card process, user authentication or system notifications in to small components that can be built independent of each other and then reassembled to form new and dynamic services that are loosely coupled and agile in that they can change as a business grows.Another key pitfall of companies doing SOA is the fact that they let vendors drive their architecture. Why do companies do this? Vendors’ do not hold your company’s success as their top priority; in fact they hold their own success as their top priority by selling you as much stuff as you are willing to buy. In my experience companies tend to strive for the maximum amount of benefits with a minimal amount of cost. Does anyone else see any conflicts between this and the driving force behind vendors.Mike Kavis recommends in an article written in CIO.com that companies need to figure out what they need before they talk to a vendor or at least have some idea of what they need. It is important to thoroughly evaluate each vendor and watch them perform a live demo of their system so that you as the company fully understand what kind of product or service the vendor is actually offering. In addition, do research on each vendor that you are considering, check out blog posts, online reviews, and any information you can find on the vendor through various search engines.Finally he recommends companies to verify any recommendations supplied by a vendor. From personal experience this is very important. I can remember when the company I worked for purchased a $200,000 add-on to their phone system that never actually worked as it was intended. In fact, just after my departure from the company started the process of attempting to get their money back from the vendor. This potentially could have been avoided if the company had done the research before selecting this vendor to ensure that their product and vendor would live up to their claims. I know that some SOA vendor offer free training regarding SOA because they know that there are a lot of misconceptions about the topic. Superficially this is a great thing for companies to take part in especially if the company is starting to implement SOA architecture and are still unsure about some topics or are looking for some guidance regarding the topic. However beware that some companies will focus on their product line only regarding the training. As an example, InfoWorld.com claims that companies providing deep seminars disguised as training, focusing more about ESBs and SOA governance technology, and less on how to approach and solve the architectural issues of the attendees.In short, it is important to remember that we as software professionals are responsible for guiding a business’s technology sections should be well informed and fully understand any new concepts that may be considered for implementation. As I have demonstrated already a company that has a few web services does not mean that they are doing SOA.  Additionally, we must not let the new buzz word of the day drive our technology, but instead our technology decisions should be driven from research and proven experience. Finally, it is important to rely on vendors when necessary, however, always take what they say with a grain of salt while cross checking any claims that they may make because we have to live with the aftermath of a system after the vendors are gone.   References: Barry, D. K. (2011). Service-oriented architecture (SOA) definition. Retrieved 12 12, 2011, from Service-Architecture.com: http://www.service-architecture.com/web-services/articles/service-oriented_architecture_soa_definition.html Connell, B. (2003, 9). service-oriented architecture (SOA). Retrieved 12 12, 2011, from SearchSOA: http://searchsoa.techtarget.com/definition/service-oriented-architecture Erl, T. (2011, 12 12). Service-Oriented Architecture. Retrieved 12 12, 2011, from WhatIsSOA: http://www.whatissoa.com/p10.php InfoWorld. (2008, 6 1). Should you get your SOA knowledge from SOA vendors? . Retrieved 12 12, 2011, from InfoWorld.com: http://www.infoworld.com/d/architecture/should-you-get-your-soa-knowledge-soa-vendors-453 Kavis, M. (2008, 6 18). Top 10 Reasons Why People are Making SOA Fail. Retrieved 12 13, 2011, from CIO.com: http://www.cio.com/article/438413/Top_10_Reasons_Why_People_are_Making_SOA_Fail?page=5&taxonomyId=3016  

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  • Why JSF Matters (to You)

    - by reza_rahman
          "Those who have knowledge, don’t predict. Those who predict, don’t have knowledge."                                                                                                    – Lao Tzu You may have noticed Thoughtworks recently crowned the likes AngularJS, etc imminent successors to server-side web frameworks. They apparently also deemed it necessary to single out JSF for righteous scorn. I have to say as I was reading the analysis I couldn't help but remember they also promptly jumped on the Ruby, Rails, Clojure, etc bandwagon a good few years ago seemingly similarly crowing these dynamic languages imminent successors to Java. I remember thinking then as I do now whether the folks at Thoughtworks are really that much smarter than me or if they are simply more prone to the Hipster buzz of the day. I'll let you make the final call on that one. I also noticed mention of "J2EE" in the context of JSF and had to wonder how up-to-date or knowledgeable the person writing the analysis actually was given that the term was basically retired almost a decade ago. There's one thing that I am absolutely sure about though - as a long time pretty happy user of JSF, I had no choice but to speak up on what I believe JSF offers. If you feel the same way, I would encourage you to support the team behind JSF whose hard work you may have benefited from over the years. True to his outspoken character PrimeFaces lead Cagatay Civici certainly did not mince words making the case for the JSF ecosystem - his excellent write-up is well worth a read. He specifically pointed out the practical problems in going whole hog with bare metal JavaScript, CSS, HTML for many development teams. I'll admit I had to smile when I read his closing sentence as well as the rather cheerful comments to the post from actual current JSF/PrimeFaces users that are apparently supposed to be on a gloomy death march. In a similar vein, OmniFaces developer Arjan Tijms did a great job pointing out the fact that despite the extremely competitive server-side Java Web UI space, JSF seems to manage to always consistently come out in either the number one or number two spot over many years and many data sources - do give his well-written message in the JAX-RS user forum a careful read. I don't think it's really reasonable to expect this to be the case for so many years if JSF was not at least a capable if not outstanding technology. If fact if you've ever wondered, Oracle itself is one of the largest JSF users on the planet. As Oracle's Shay Shmeltzer explains in a recent JSF Central interview, many of Oracle's strategic products such as ADF, ADF Mobile and Fusion Applications itself is built on JSF. There are well over 3,000 active developers working on these codebases. I don't think anyone can think of a more compelling reason to make sure that a technology is as effective as possible for practical development under real world conditions. Standing on the shoulders of the above giants, I feel like I can be pretty brief in making my own case for JSF: JSF is a powerful abstraction that brings the original Smalltalk MVC pattern to web development. This means cutting down boilerplate code to the bare minimum such that you really can think of just writing your view markup and then simply wire up some properties and event handlers on a POJO. The best way to see what this really means is to compare JSF code for a pretty small case to other approaches. You should then multiply the additional work for the typical enterprise project to try to understand what the productivity trade-offs are. This is reason alone for me to personally never take any other approach seriously as my primary web UI solution unless it can match the sheer productivity of JSF. Thanks to JSF's focus on components from the ground-up JSF has an extremely strong ecosystem that includes projects like PrimeFaces, RichFaces, OmniFaces, ICEFaces and of course ADF Faces/Mobile. These component libraries taken together constitute perhaps the largest widget set ever developed and optimized for a single web UI technology. To begin to grasp what this really means, just briefly browse the excellent PrimeFaces showcase and think about the fact that you can readily use the widgets on that showcase by just using some simple markup and knowing near to nothing about AJAX, JavaScript or CSS. JSF has the fair and legitimate advantage of being an open vendor neutral standard. This means that no single company, individual or insular clique controls JSF - openness, transparency, accountability, plurality, collaboration and inclusiveness is virtually guaranteed by the standards process itself. You have the option to choose between compatible implementations, escape any form of lock-in or even create your own compatible implementation! As you might gather from the quote at the top of the post, I am not a fan of crystal ball gazing and certainly don't want to engage in it myself. Who knows? However far-fetched it may seem maybe AngularJS is the only future we all have after all. If that is the case, so be it. Unlike what you might have been told, Java EE is about choice at heart and it can certainly work extremely well as a back-end for AngularJS. Likewise, you are also most certainly not limited to just JSF for working with Java EE - you have a rich set of choices like Struts 2, Vaadin, Errai, VRaptor 4, Wicket or perhaps even the new action-oriented web framework being considered for Java EE 8 based on the work in Jersey MVC... Please note that any views expressed here are my own only and certainly does not reflect the position of Oracle as a company.

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  • Controlling the Sizing of the af:messages Dialog

    - by Duncan Mills
    Over the last day or so a small change in behaviour between 11.1.2.n releases of ADF and earlier versions has come to my attention. This has concerned the default sizing of the dialog that the framework automatically generates to handle the display of JSF messages being handled by the <af:messages> component. Unlike a normal popup, you don't have a physical <af:dialog> or <af:window> to set the sizing on in your page definition, so you're at the mercy of what the framework provides. In this case the framework now defines a fixed 250x250 pixel content area dialog for these messages, which can look a bit weird if the message is either very short, or very long. Unfortunately this is not something that you can control through the skin, instead you have to be a little more creative. Here's the solution I've come up with.  Unfortunately, I've not found a supportable way to reset the dialog so as to say  just size yourself based on your contents, it is actually possible to do this by tweaking the correct DOM objects, but I wanted to start with a mostly supportable solution that only uses the best practice of working through the ADF client side APIs. The Technique The basic approach I've taken is really very simple.  The af:messages dialog is just a normal richDialog object, it just happens to be one that is pre-defined for you with a particular known name "msgDlg" (which hopefully won't change). Knowing this, you can call the accepted APIs to control the content width and height of that dialog, as our meerkat friends would say, "simples" 1 The JavaScript For this example I've defined three JavaScript functions.   The first does all the hard work and is designed to be called from server side Java or from a page load event to set the default. The second is a utility function used by the first to validate the values you're about to use for height and width. The final function is one that can be called from the page load event to set an initial default sizing if that's all you need to do. Function resizeDefaultMessageDialog() /**  * Function that actually resets the default message dialog sizing.  * Note that the width and height supplied define the content area  * So the actual physical dialog size will be larger to account for  * the chrome containing the header / footer etc.  * @param docId Faces component id of the document  * @param contentWidth - new content width you need  * @param contentHeight - new content height  */ function resizeDefaultMessageDialog(docId, contentWidth, contentHeight) {   // Warning this value may change from release to release   var defMDName = "::msgDlg";   //Find the default messages dialog   msgDialogComponent = AdfPage.PAGE.findComponentByAbsoluteId(docId + defMDName); // In your version add a check here to ensure we've found the right object!   // Check the new width is supplied and is a positive number, if so apply it.   if (dimensionIsValid(contentWidth)){       msgDialogComponent.setContentWidth(contentWidth);   }   // Check the new height is supplied and is a positive number, if so apply it.   if (dimensionIsValid(contentHeight)){       msgDialogComponent.setContentHeight(contentHeight);   } }  Function dimensionIsValid()  /**  * Simple function to check that sensible numeric values are   * being proposed for a dimension  * @param sampleDimension   * @return booolean  */ function dimensionIsValid(sampleDimension){     return (!isNaN(sampleDimension) && sampleDimension > 0); } Function  initializeDefaultMessageDialogSize() /**  * This function will re-define the default sizing applied by the framework   * in 11.1.2.n versions  * It is designed to be called with the document onLoad event  */ function initializeDefaultMessageDialogSize(loadEvent){   //get the configuration information   var documentId = loadEvent.getSource().getProperty('documentId');   var newWidth = loadEvent.getSource().getProperty('defaultMessageDialogContentWidth');   var newHeight = loadEvent.getSource().getProperty('defaultMessageDialogContentHeight');   resizeDefaultMessageDialog(documentId, newWidth, newHeight); } Wiring in the Functions As usual, the first thing we need to do when using JavaScript with ADF is to define an af:resource  in the document metaContainer facet <af:document>   ....     <f:facet name="metaContainer">     <af:resource type="javascript" source="/resources/js/hackMessagedDialog.js"/>    </f:facet> </af:document> This makes the script functions available to call.  Next if you want to use the option of defining an initial default size for the dialog you use a combination of <af:clientListener> and <af:clientAttribute> tags like this. <af:document title="MyApp" id="doc1">   <af:clientListener method="initializeDefaultMessageDialogSize" type="load"/>   <af:clientAttribute name="documentId" value="doc1"/>   <af:clientAttribute name="defaultMessageDialogContentWidth" value="400"/>   <af:clientAttribute name="defaultMessageDialogContentHeight" value="150"/>  ...   Just in Time Dialog Sizing  So  what happens if you have a variety of messages that you might add and in some cases you need a small dialog and an other cases a large one? Well in that case you can re-size these dialogs just before you submit the message. Here's some example Java code: FacesContext ctx = FacesContext.getCurrentInstance();          //reset the default dialog size for this message ExtendedRenderKitService service =              Service.getRenderKitService(ctx, ExtendedRenderKitService.class); service.addScript(ctx, "resizeDefaultMessageDialog('doc1',100,50);");          FacesMessage msg = new FacesMessage("Short message"); msg.setSeverity(FacesMessage.SEVERITY_ERROR); ctx.addMessage(null, msg);  So there you have it. This technique should, at least, allow you to control the dialog sizing just enough to stop really objectionable whitespace or scrollbars. 1 Don't worry if you don't get the reference, lest's just say my kids watch too many adverts.

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  • Configure PERL DBI and DBD in Linux

    - by Balualways
    I am new to Perl and I work in a Linux OEL 5x server. I am trying to configure the Perl DB modules for Oracle connectivity (DBD and DBI modules). Can anyone help me out in the installation procedure? I had tried CPAN didn't really worked out. Any help would be appreciated. I am not quite sure I need to initialize any variables other than $LD_LIBRARY_PATH and $ORACLE_HOME These are my observations: ISSUE:: I am getting the following issue while using the DBI module to connect to Oracle: install_driver(Oracle) failed: Can't locate loadable object for module DBD::Oracle in @INC (@INC contains: /usr/lib64/perl5/site_perl/5.8.8/x86_64-linux-thread-multi /usr/lib/perl5/site_perl/5.8.8 /usr/lib/perl5/site_perl /usr/lib64/perl5/vendor_perl/5.8.8/x86_64-linux-thread-multi /usr/lib/perl5/vendor_perl/5.8.8 /usr/lib/perl5/vendor_perl /usr/lib64/perl5/5.8.8/x86_64-linux-thread-multi /usr/lib/perl5/5.8.8 .) at (eval 3) line 3 Compilation failed in require at (eval 3) line 3. Perhaps a module that DBD::Oracle requires hasn't been fully installed at connectdb.pl line 57 I had installed the DBD for oracle from /usr/lib64/perl5/5.8.8/x86_64-linux-thread-multi/DBD/DBD-Oracle-1.50 Could you please take a look into the steps and correct me if I am wrong: Observations: $ echo $LD_LIBRARY_PATH /opt/CA/UnicenterAutoSysJM/autosys/lib:/opt/CA/SharedComponents/Csam/SockAdapter/lib:/opt/CA/SharedComponents/ETPKI/lib:/opt/CA/CAlib $ echo $ORACLE_HOME /usr/local/oracle/ORA This is how I tried to install the DBD module: Download the file DBD 1.50 for Oracle Copy to /usr/lib64/perl5/5.8.8/x86_64-linux-thread-multi/DBD Untar and Makefile.PL . Message: Using DBI 1.52 (for perl 5.008008 on x86_64-linux-thread-multi) installed in /usr/lib64/perl5/vendor_perl/5.8.8/x86_64-linux-thread-multi/auto/DBI/ Configuring DBD::Oracle for perl 5.008008 on linux (x86_64-linux-thread-multi) Remember to actually *READ* the README file! Especially if you have any problems. Installing on a linux, Ver#2.6 Using Oracle in /opt/oracle/product/10.2 DEFINE _SQLPLUS_RELEASE = "1002000400" (CHAR) Oracle version 10.2.0.4 (10.2) Found /opt/oracle/product/10.2/rdbms/demo/demo_rdbms.mk Found /opt/oracle/product/10.2/rdbms/demo/demo_rdbms64.mk Found /opt/oracle/product/10.2/rdbms/lib/ins_rdbms.mk Using /opt/oracle/product/10.2/rdbms/demo/demo_rdbms.mk Your LD_LIBRARY_PATH env var is set to '/usr/local/oracle/ORA/lib:/usr/dt/lib:/usr/openwin/lib:/usr/local/oracle/ORA/ows/cartx/wodbc/1.0/util/lib:/usr/local/oracle/ORA/lib:/usr/local/sybase/OCS-12_0/lib:/usr/local/sybase/lib:/home/oracle/jdbc/jdbcoci73/lib:./' WARNING: Your LD_LIBRARY_PATH env var doesn't include '/opt/oracle/product/10.2/lib' but probably needs to. Reading /opt/oracle/product/10.2/rdbms/demo/demo_rdbms.mk Reading /usr/local/oracle/ORA/rdbms/lib/env_rdbms.mk Attempting to discover Oracle OCI build rules sh: make: command not found by executing: [make -f /opt/oracle/product/10.2/rdbms/demo/demo_rdbms.mk build ECHODO=echo ECHO=echo GENCLNTSH='echo genclntsh' CC=true OPTIMIZE= CCFLAGS= EXE=DBD_ORA_EXE OBJS=DBD_ORA_OBJ.o] WARNING: Oracle build rule discovery failed (32512) Add path to make command into your PATH environment variable. Oracle oci build prolog: [sh: make: command not found] Oracle oci build command: [] WARNING: Unable to interpret Oracle build commands from /opt/oracle/product/10.2/rdbms/demo/demo_rdbms.mk. (Will continue by using fallback approach.) Please report this to [email protected]. See README for what to include. Found header files in /opt/oracle/product/10.2/rdbms/public. client_version=10.2 DEFINE= -Wall -Wno-comment -DUTF8_SUPPORT -DORA_OCI_VERSION=\"10.2.0.4\" -DORA_OCI_102 Checking for functioning wait.ph System: perl5.008008 linux ca-build9.us.oracle.com 2.6.20-1.3002.fc6xen #1 smp thu apr 30 18:08:39 pdt 2009 x86_64 x86_64 x86_64 gnulinux Compiler: gcc -O2 -g -pipe -Wall -Wp,-D_FORTIFY_SOURCE=2 -fexceptions -fstack-protector --param=ssp-buffer-size=4 -m64 -mtune=generic -D_REENTRANT -D_GNU_SOURCE -fno-strict-aliasing -pipe -Wdeclaration-after-statement -I/usr/local/include -D_LARGEFILE_SOURCE -D_FILE_OFFSET_BITS=64 -I/usr/include/gdbm Linker: not found Sysliblist: -ldl -lm -lpthread -lnsl -lirc Oracle makefiles would have used these definitions but we override them: CC: cc CFLAGS: $(GFLAG) $(OPTIMIZE) $(CDEBUG) $(CCFLAGS) $(PFLAGS)\ $(SHARED_CFLAG) $(USRFLAGS) [$(GFLAG) -O3 $(CDEBUG) -m32 $(TRIGRAPHS_CCFLAGS) -fPIC -I/usr/local/oracle/ORA/rdbms/demo -I/usr/local/oracle/ORA/rdbms/public -I/usr/local/oracle/ORA/plsql/public -I/usr/local/oracle/ORA/network/public -DLINUX -D_GNU_SOURCE -D_LARGEFILE64_SOURCE=1 -D_LARGEFILE_SOURCE=1 -DSLTS_ENABLE -DSLMXMX_ENABLE -D_REENTRANT -DNS_THREADS -fno-strict-aliasing $(LPFLAGS) $(USRFLAGS)] build: $(CC) $(ORALIBPATH) -o $(EXE) $(OBJS) $(OCISHAREDLIBS) [ cc -L$(LIBHOME) -L/usr/local/oracle/ORA/rdbms/lib/ -o $(EXE) $(OBJS) -lclntsh $(EXPDLIBS) $(EXOSLIBS) -ldl -lm -lpthread -lnsl -lirc -ldl -lm $(USRLIBS) -lpthread] LDFLAGS: $(LDFLAGS32) [-m32 -o $@ -L/usr/local/oracle/ORA/rdbms//lib32/ -L/usr/local/oracle/ORA/lib32/ -L/usr/local/oracle/ORA/lib32/stubs/] Linking with /usr/local/oracle/ORA/rdbms/lib/defopt.o -lclntsh -ldl -lm -lpthread -lnsl -lirc -ldl -lm -lpthread [from $(DEF_OPT) $(OCISHAREDLIBS)] Checking if your kit is complete... Looks good LD_RUN_PATH=/usr/local/oracle/ORA/lib Using DBD::Oracle 1.50. Using DBD::Oracle 1.50. Using DBI 1.52 (for perl 5.008008 on x86_64-linux-thread-multi) installed in /usr/lib64/perl5/vendor_perl/5.8.8/x86_64-linux-thread-multi/auto/DBI/ Writing Makefile for DBD::Oracle Writing MYMETA.yml and MYMETA.json *** If you have problems... read all the log printed above, and the README and README.help.txt files. (Of course, you have read README by now anyway, haven't you?)

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  • Two network interfaces and two IP addresses on the same subnet in Linux

    - by Scott Duckworth
    I recently ran into a situation where I needed two IP addresses on the same subnet assigned to one Linux host so that we could run two SSL/TLS sites. My first approach was to use IP aliasing, e.g. using eth0:0, eth0:1, etc, but our network admins have some fairly strict settings in place for security that squashed this idea: They use DHCP snooping and normally don't allow static IP addresses. Static addressing is accomplished by using static DHCP entries, so the same MAC address always gets the same IP assignment. This feature can be disabled per switchport if you ask and you have a reason for it (thankfully I have a good relationship with the network guys and this isn't hard to do). With the DHCP snooping disabled on the switchport, they had to put in a rule on the switch that said MAC address X is allowed to have IP address Y. Unfortunately this had the side effect of also saying that MAC address X is ONLY allowed to have IP address Y. IP aliasing required that MAC address X was assigned two IP addresses, so this didn't work. There may have been a way around these issues on the switch configuration, but in an attempt to preserve good relations with the network admins I tried to find another way. Having two network interfaces seemed like the next logical step. Thankfully this Linux system is a virtual machine, so I was able to easily add a second network interface (without rebooting, I might add - pretty cool). A few keystrokes later I had two network interfaces up and running and both pulled IP addresses from DHCP. But then the problem came in: the network admins could see (on the switch) the ARP entry for both interfaces, but only the first network interface that I brought up would respond to pings or any sort of TCP or UDP traffic. After lots of digging and poking, here's what I came up with. It seems to work, but it also seems to be a lot of work for something that seems like it should be simple. Any alternate ideas out there? Step 1: Enable ARP filtering on all interfaces: # sysctl -w net.ipv4.conf.all.arp_filter=1 # echo "net.ipv4.conf.all.arp_filter = 1" >> /etc/sysctl.conf From the file networking/ip-sysctl.txt in the Linux kernel docs: arp_filter - BOOLEAN 1 - Allows you to have multiple network interfaces on the same subnet, and have the ARPs for each interface be answered based on whether or not the kernel would route a packet from the ARP'd IP out that interface (therefore you must use source based routing for this to work). In other words it allows control of which cards (usually 1) will respond to an arp request. 0 - (default) The kernel can respond to arp requests with addresses from other interfaces. This may seem wrong but it usually makes sense, because it increases the chance of successful communication. IP addresses are owned by the complete host on Linux, not by particular interfaces. Only for more complex setups like load- balancing, does this behaviour cause problems. arp_filter for the interface will be enabled if at least one of conf/{all,interface}/arp_filter is set to TRUE, it will be disabled otherwise Step 2: Implement source-based routing I basically just followed directions from http://lartc.org/howto/lartc.rpdb.multiple-links.html, although that page was written with a different goal in mind (dealing with two ISPs). Assume that the subnet is 10.0.0.0/24, the gateway is 10.0.0.1, the IP address for eth0 is 10.0.0.100, and the IP address for eth1 is 10.0.0.101. Define two new routing tables named eth0 and eth1 in /etc/iproute2/rt_tables: ... top of file omitted ... 1 eth0 2 eth1 Define the routes for these two tables: # ip route add default via 10.0.0.1 table eth0 # ip route add default via 10.0.0.1 table eth1 # ip route add 10.0.0.0/24 dev eth0 src 10.0.0.100 table eth0 # ip route add 10.0.0.0/24 dev eth1 src 10.0.0.101 table eth1 Define the rules for when to use the new routing tables: # ip rule add from 10.0.0.100 table eth0 # ip rule add from 10.0.0.101 table eth1 The main routing table was already taken care of by DHCP (and it's not even clear that its strictly necessary in this case), but it basically equates to this: # ip route add default via 10.0.0.1 dev eth0 # ip route add 130.127.48.0/23 dev eth0 src 10.0.0.100 # ip route add 130.127.48.0/23 dev eth1 src 10.0.0.101 And voila! Everything seems to work just fine. Sending pings to both IP addresses works fine. Sending pings from this system to other systems and forcing the ping to use a specific interface works fine (ping -I eth0 10.0.0.1, ping -I eth1 10.0.0.1). And most importantly, all TCP and UDP traffic to/from either IP address works as expected. So again, my question is: is there a better way to do this? This seems like a lot of work for a seemingly simple problem.

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  • rm on a directory with millions of files

    - by BMDan
    Background: physical server, about two years old, 7200-RPM SATA drives connected to a 3Ware RAID card, ext3 FS mounted noatime and data=ordered, not under crazy load, kernel 2.6.18-92.1.22.el5, uptime 545 days. Directory doesn't contain any subdirectories, just millions of small (~100 byte) files, with some larger (a few KB) ones. We have a server that has gone a bit cuckoo over the course of the last few months, but we only noticed it the other day when it started being unable to write to a directory due to it containing too many files. Specifically, it started throwing this error in /var/log/messages: ext3_dx_add_entry: Directory index full! The disk in question has plenty of inodes remaining: Filesystem Inodes IUsed IFree IUse% Mounted on /dev/sda3 60719104 3465660 57253444 6% / So I'm guessing that means we hit the limit of how many entries can be in the directory file itself. No idea how many files that would be, but it can't be more, as you can see, than three million or so. Not that that's good, mind you! But that's part one of my question: exactly what is that upper limit? Is it tunable? Before I get yelled at--I want to tune it down; this enormous directory caused all sorts of issues. Anyway, we tracked down the issue in the code that was generating all of those files, and we've corrected it. Now I'm stuck with deleting the directory. A few options here: rm -rf (dir)I tried this first. I gave up and killed it after it had run for a day and a half without any discernible impact. unlink(2) on the directory: Definitely worth consideration, but the question is whether it'd be faster to delete the files inside the directory via fsck than to delete via unlink(2). That is, one way or another, I've got to mark those inodes as unused. This assumes, of course, that I can tell fsck not to drop entries to the files in /lost+found; otherwise, I've just moved my problem. In addition to all the other concerns, after reading about this a bit more, it turns out I'd probably have to call some internal FS functions, as none of the unlink(2) variants I can find would allow me to just blithely delete a directory with entries in it. Pooh. while [ true ]; do ls -Uf | head -n 10000 | xargs rm -f 2/dev/null; done ) This is actually the shortened version; the real one I'm running, which just adds some progress-reporting and a clean stop when we run out of files to delete, is: export i=0; time ( while [ true ]; do ls -Uf | head -n 3 | grep -qF '.png' || break; ls -Uf | head -n 10000 | xargs rm -f 2/dev/null; export i=$(($i+10000)); echo "$i..."; done ) This seems to be working rather well. As I write this, it's deleted 260,000 files in the past thirty minutes or so. Now, for the questions: As mentioned above, is the per-directory entry limit tunable? Why did it take "real 7m9.561s / user 0m0.001s / sys 0m0.001s" to delete a single file which was the first one in the list returned by "ls -U", and it took perhaps ten minutes to delete the first 10,000 entries with the command in #3, but now it's hauling along quite happily? For that matter, it deleted 260,000 in about thirty minutes, but it's now taken another fifteen minutes to delete 60,000 more. Why the huge swings in speed? Is there a better way to do this sort of thing? Not store millions of files in a directory; I know that's silly, and it wouldn't have happened on my watch. Googling the problem and looking through SF and SO offers a lot of variations on "find" that obviously have the wrong idea; it's not going to be faster than my approach for several self-evident reasons. But does the delete-via-fsck idea have any legs? Or something else entirely? I'm eager to hear out-of-the-box (or inside-the-not-well-known-box) thinking. Thanks for reading the small novel; feel free to ask questions and I'll be sure to respond. I'll also update the question with the final number of files and how long the delete script ran once I have that. Final script output!: 2970000... 2980000... 2990000... 3000000... 3010000... real 253m59.331s user 0m6.061s sys 5m4.019s So, three million files deleted in a bit over four hours.

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  • Performance required to improve Windows Experience Index?

    - by Ian Boyd
    Is there a guide on the metrics required to obtain a certain Windows Experience Index? A Microsoft guy said in January 2009: On the matter of transparency, it is indeed our plan to disclose in great detail how the scores are calculated, what the tests attempt to measure, why, and how they map to realistic scenarios and usage patterns. Has that amount of transparency happened? Is there a technet article somewhere? If my score was limited by my Memory subscore of 5.9. A nieve person would suggest: Buy a faster RAM Which is wrong of course. From the Windows help: If your computer has a 64-bit central processing unit (CPU) and 4 gigabytes (GB) or less random access memory (RAM), then the Memory (RAM) subscore for your computer will have a maximum of 5.9. You can buy the fastest, overclocked, liquid-cooled, DDR5 RAM on the planet; you'll still have a maximum Memory subscore of 5.9. So in general the knee-jerk advice "buy better stuff" is not helpful. What i am looking for is attributes required to achieve a certain score, or move beyond a current limitation. The information i've been able to compile so far, chiefly from 3 Windows blog entries, and an article: Memory subscore Score Conditions ======= ================================ 1.0 < 256 MB 2.0 < 500 MB 2.9 <= 512 MB 3.5 < 704 MB 3.9 < 944 MB 4.5 <= 1.5 GB 5.9 < 4.0GB-64MB on a 64-bit OS Windows Vista highest score 7.9 Windows 7 highest score Graphics Subscore Score Conditions ======= ====================== 1.0 doesn't support DX9 1.9 doesn't support WDDM 4.9 does not support Pixel Shader 3.0 5.9 doesn't support DX10 or WDDM1.1 Windows Vista highest score 7.9 Windows 7 highest score Gaming graphics subscore Score Result ======= ============================= 1.0 doesn't support D3D 2.0 supports D3D9, DX9 and WDDM 5.9 doesn't support DX10 or WDDM1.1 Windows Vista highest score 6.0-6.9 good framerates (e.g. 40-50fps) at normal resoltuions (e.g. 1280x1024) 7.0-7.9 even higher framerates at even higher resolutions 7.9 Windows 7 highest score Processor subscore Score Conditions ======= ========================================================================== 5.9 Windows Vista highest score 6.0-6.9 many quad core processors will be able to score in the high 6 low 7 ranges 7.0+ many quad core processors will be able to score in the high 6 low 7 ranges 7.9 8-core systems will be able to approach 8.9 Windows 7 highest score Primary hard disk subscore (note) Score Conditions ======= ======================================== 1.9 Limit for pathological drives that stop responding when pending writes 2.0 Limit for pathological drives that stop responding when pending writes 2.9 Limit for pathological drives that stop responding when pending writes 3.0 Limit for pathological drives that stop responding when pending writes 5.9 highest you're likely to see without SSD Windows Vista highest score 7.9 Windows 7 highest score Bonus Chatter You can find your WEI detailed test results in: C:\Windows\Performance\WinSAT\DataStore e.g. 2011-11-06 01.00.19.482 Disk.Assessment (Recent).WinSAT.xml <WinSAT> <WinSPR> <DiskScore>5.9</DiskScore> </WinSPR> <Metrics> <DiskMetrics> <AvgThroughput units="MB/s" score="6.4" ioSize="65536" kind="Sequential Read">89.95188</AvgThroughput> <AvgThroughput units="MB/s" score="4.0" ioSize="16384" kind="Random Read">1.58000</AvgThroughput> <Responsiveness Reason="UnableToAssess" Kind="Cap">TRUE</Responsiveness> </DiskMetrics> </Metrics> </WinSAT> Pre-emptive snarky comment: "WEI is useless, it has no relation to reality" Fine, how do i increase my hard-drive's random I/O throughput? Update - Amount of memory limits rating Some people don't believe Microsoft's statement that having less than 4GB of RAM on a 64-bit edition of Windows doesn't limit the rating to 5.9: And from xxx.Formal.Assessment (Recent).WinSAT.xml: <WinSPR> <LimitsApplied> <MemoryScore> <LimitApplied Friendly="Physical memory available to the OS is less than 4.0GB-64MB on a 64-bit OS : limit mem score to 5.9" Relation="LT">4227858432</LimitApplied> </MemoryScore> </LimitsApplied> </WinSPR> References Windows Vista Team Blog: Windows Experience Index: An In-Depth Look Understand and improve your computer's performance in Windows Vista Engineering Windows 7 Blog: Engineering the Windows 7 “Windows Experience Index”

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  • Mount TMPFS instead of ro /dev

    - by schiggn
    I am working on a ARM-Based embedded system with a custom Debian Linux based on kernel 2.6.31. In the final system, the Root file system is stored as squashfs on flash. Now, the folder /dev is created by udev, but since there is no hot plugging functionality needed and booting time is critical, I wanted to delete udev and "hard code" the /dev folder (read here, page 5). because i still need to change parameters of the devices (with ioctl /sysfs) this does not work for me in this case. so i thought of mounting a tmpfs on /dev and change the parameters there. is this possible? and how to do best? my approach would be: delete /dev from RFS create tar containing basic devices mount tmpfs /dev untar tar-file into /dev change parameters Could this work? Do you see any problems? I found out, that you can mount on top of already mounted mount point, is it somehow possible just to take data with while mounting the new file system? if so that would be very convenient! Thanks Update: I just tried that out, but I'm stuck at a certain point. I packed all my devices into devices.tar, packed it into /usr of my squashfs and added the following lines to mountkernfs.sh, which is executed right after INIT. #mount /dev on tmpfs echo -n "Mounting /dev on tmpfs..." mount -o size=5M,mode=0755 -t tmpfs tmpfs /dev mknod -m 600 /dev/console c 5 1 mknod -m 600 /dev/null c 1 3 echo "done." echo -n "Populating /dev..." tar -xf /usr/devices.tar -C /dev echo "done." This works fine on the version over NFS, if I place printf's in the code, I can see it executing, if I comment out the extracting part, its complaining about missing devices. Booting OK mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 IP-Config: Unable to set interface netmask (-22). Looking up port of RPC 100003/2 on 192.168.1.234 Looking up port of RPC 100005/1 on 192.168.1.234 VFS: Mounted root (nfs filesystem) on device 0:14. Freeing init memory: 136K INIT: version 2.86 booting Mounting /dev on tmpfs...done. Populating /dev...done. Initializing /var...done. Setting the system clock. System Clock set to: Thu Sep 13 11:26:23 UTC 2012. INIT: Entering runlevel: 2 UBI: attaching mtd8 to ubi0 Commenting out the extraction of the tar mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 IP-Config: Unable to set interface netmask (-22). Looking up port of RPC 100003/2 on 192.168.1.234 Looking up port of RPC 100005/1 on 192.168.1.234 VFS: Mounted root (nfs filesystem) on device 0:14. Freeing init memory: 136K INIT: version 2.86 booting Mounting /dev on tmpfs...done. Populating /dev...done. Initializing /var...done. Setting the system clock. Cannot access the Hardware Clock via any known method. Use the --debug option to see the details of our search for an access method. Unable to set System Clock to: Thu Sep 13 12:24:00 UTC 2012 ... (warning). INIT: Entering runlevel: 2 libubi: error!: cannot open "/dev/ubi_ctrl" So far so good. But if I pack the whole story into a squashfs and boot from there, it is acting strange. It's telling me while booting that it is unable to open an initial console and its throwing errors on mounting the UBIFS devices, but finally provides a login anyway. Over that my echo's are not executed. If I then log in, /dev is mounted as TMPFS as desired and all the devices reside inside. When I redo the "mount" command to mount the UBIFS partitions it is executed whitout problem and useable. From squashfs VFS: Mounted root (squashfs filesystem) readonly on device 31:15. Freeing init memory: 136K Warning: unable to open an initial console. mmc0: new high speed SDHC card at address 0007 mmcblk0: mmc0:0007 SD04G 3.67 GiB mmcblk0: p1 UBIFS error (pid 484): ubifs_get_sb: cannot open "ubi1_0", error -19 Additionally, a part of the rest of the bootscripts is still exexuted, but not all of them. Does anyone has a clue why? Other question, is 5MB enough/too much for /dev?

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  • Clearing C#'s WebBrowser control's cookies for all sites WITHOUT clearing for IE itself

    - by Helgi Hrafn Gunnarsson
    Hail StackOverflow! The short version of what I'm trying to do is in the title. Here's the long version. I have a bit of a complex problem which I'm sure I will receive a lot of guesses as a response to. In order to keep the well-intended but unfortunately useless guesses to a minimum, let me first mention that the solution to this problem is not simple, so simple suggestions will unfortunately not help at all, even though I appreciate the effort. C#'s WebBrowser component is fundamentally IE itself so solutions with any sorts of caveats will almost certainly not work. I need to do exactly what I'm trying to do, and even a seemingly minor caveat will defeat the purpose completely. At the risk of sounding arrogant, I need assistance from someone who really has in-depth knowledge about C#'s WebBrowser and/or WinInet and/or how to communicate with Windows's underlying system from C#... or how to encapsulate C++ code in C#. That said, I don't expect anyone to do this for me, and I've found some promising hints which are explained later in this question. But first... what I'm trying to achieve is this. I have a Windows.Forms component which contains a WebBrowser control. This control needs to: Clear ALL cookies for ALL websites. Visit several websites, one after another, and record cookies and handle them correctly. This part works fine already so I don't have any problems with this. Rinse and repeat... theoretically forever. Now, here's the real problem. I need to clear all those cookies (for any and all sites), but only for the WebBrowser control itself and NOT the cookies which IE proper uses. What's fundamentally wrong with this approach is of course the fact that C#'s WebBrowser control is IE. But I'm a stubborn young man and I insist on it being possible, or else! ;) Here's where I'm stuck at the moment. It is quite simply impossible to clear all cookies for the WebBrowser control programmatically through C# alone. One must use DllImport and all the crazy stuff that comes with it. This chunk works fine for that purpose: [DllImport("wininet.dll", SetLastError = true)] private static extern bool InternetSetOption(IntPtr hInternet, int dwOption, IntPtr lpBuffer, int lpdwBufferLength); And then, in the function that actually does the clearing of the cookies: InternetSetOption(IntPtr.Zero, INTERNET_OPTION_END_BROWSER_SESSION, IntPtr.Zero, 0); Then all the cookies get cleared and as such, I'm happy. The program works exactly as intended, aside from the fact that it also clears IE's cookies, which must not be allowed to happen. The problem is that this also clears the cookies for IE proper, and I can't have that happen. From one fellow StackOverflower (if that's a word), Sheng Jiang proposed this to a different problem in a comment, but didn't elaborate further: "If you want to isolate your application's cookies you need to override the Cache directory registry setting via IDocHostUIHandler2::GetOverrideKeyPath" I've looked around the internet for IDocHostUIHandler2 and GetOverrideKeyPath, but I've got no idea of how to use them from C# to isolate cookies to my WebBrowser control. My experience with the Windows registry is limited to RegEdit (so I understand that it's a tree structure with different data types but that's about it... I have no in-depth knowledge of the registry's relationship with IE, for example). Here's what I dug up on MSDN: IDocHostUIHandler2 docs: http://msdn.microsoft.com/en-us/library/aa753275%28VS.85%29.aspx GetOverrideKeyPath docs: http://msdn.microsoft.com/en-us/library/aa753274%28VS.85%29.aspx I think I know roughly what these things do, I just don't know how to use them. So, I guess that's it! Any help is greatly appreciated.

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  • C# Bind DataTable to Existing DataGridView Column Definitions

    - by Timothy
    I've been struggling with a NullReferenceException and hope someone here will be able to point me in the right direction. I'm trying to create and populate a DataTable and then show the results in a DataGridView control. The basic code follows, and Execution stops with a NullReferenceException at the point where I invoke the new UpdateResults_Delegate. Oddly enough, I can trace entries.Rows.Count successfully before I return it from QueryEventEntries, so I can at least show 1) entries is not a null reference, and 2) the DataTable contains rows of data. I know I have to be doing something wrong, but I just don't know what. private void UpdateResults(DataTable entries) { dataGridView.DataSource = entries; } private void button_Click(object sender, EventArgs e) { PerformQuery(); } private void PerformQuery() { DateTime start = new DateTime(dateTimePicker1.Value.Year, dateTimePicker1.Value.Month, dateTimePicker1.Value.Day, 0, 0, 0); DateTime stop = new DateTime(dateTimePicker2.Value.Year, dateTimePicker2.Value.Month, dateTimePicker2.Value.Day, 0, 0, 0); DataTable entries = QueryEventEntries(start, stop); UpdateResults(entries); } private DataTable QueryEventEntries(DateTime start, DateTime stop) { DataTable entries = new DataTable(); entries.Columns.AddRange(new DataColumn[] { new DataColumn("event_type", typeof(Int32)), new DataColumn("event_time", typeof(DateTime)), new DataColumn("event_detail", typeof(String))}); using (SqlConnection conn = new SqlConnection(DSN)) { using (SqlDataAdapter adapter = new SqlDataAdapter( "SELECT event_type, event_time, event_detail FROM event_log " + "WHERE event_time >= @start AND event_time <= @stop", conn)) { adapter.SelectCommand.Parameters.AddRange(new Object[] { new SqlParameter("@start", start), new SqlParameter("@stop", stop)}); adapter.Fill(entries); } } return entries; } Update I'd like to summarize and provide some additional information I've learned from the discussion here and debugging efforts since I originally posted this question. I am refactoring old code that retrieved records from a database, collected those records as an array, and then later iterated through the array to populate a DataGridView row by row. Threading was originally implemented to compensate and keep the UI responsive during the unnecessary looping. I have since stripped out Thread/Invoke; everything now occurs on the same execution thread (thank you, Sam). I am attempting to replace the slow, unwieldy approach using a DataTable which I can fill with a DataAdapter, and assign to the DataGridView through it's DataSource property (above code updated). I've iterated through the entries DataTable's rows to verify the table contains the expected data before assigning it as the DataGridView's DataSource. foreach (DataRow row in entries.Rows) { System.Diagnostics.Trace.WriteLine( String.Format("{0} {1} {2}", row[0], row[1], row[2])); } One of the column of the DataGridView is a custom DataGridViewColumn to stylize the event_type value. I apologize I didn't mention this before in the original post but I wasn't aware it was important to my problem. I have converted this column temporarily to a standard DataGridViewTextBoxColumn control and am no longer experiencing the Exception. The fields in the DataTable are appended to the list of fields that have been pre-specified in Design view of the DataGridView. The records' values are being populated in these appended fields. When the run time attempts to render the cell a null value is provided (as the value that should be rendered is done so a couple columns over). In light of this, I am re-titling and re-tagging the question. I would still appreciate it if others who have experienced this can instruct me on how to go about binding the DataTable to the existing column definitions of the DataGridView.

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  • Binding navigation property to RadGrid while using EntityDataSource control

    - by Matrix
    I'm new to Entity Framework and I got stuck in an issue while trying to bind a navigation property (foreign key reference) to a dropdownlist. I have Telerik RadGrid control which gets the data using a EntityDataSource control. Here is the model description: Applications: AppId, AppName, ServerId Servers: ServerId, ServerName The Applicaitons.ServerId is a foreign key reference to Servers.ServerId. The RadGrid lists the applications and allows the user to insert/update/delete an application. I want to show the server names as a dropdownlist in edit mode which I'm not able to. . Here is my aspx code: <telerik:RadGrid ID="gridApplications" runat="server" Skin="Sunset" AllowAutomaticInserts="True" AllowAutomaticDeletes="True" AllowPaging="True" AllowAutomaticUpdates="True" AutoGenerateColumns="False" OnItemCreated="gridApplications_ItemCreated" DataSourceID="applicationsEntityDataSource" Width="50%" OnItemInserted="gridApplications_ItemInserted" OnItemUpdated="gridApplications_ItemUpdated" OnItemDeleted="gridApplications_ItemDeleted" GridLines="None"> <MasterTableView CommandItemDisplay="Top" AutoGenerateColumns="False" DataKeyNames="AppId" DataSourceID="applicationsEntityDataSource"> <RowIndicatorColumn> <HeaderStyle Width="20px" /> </RowIndicatorColumn> <ExpandCollapseColumn> <HeaderStyle Width="20px" /> </ExpandCollapseColumn> <Columns> <telerik:GridEditCommandColumn ButtonType="ImageButton" UniqueName="EditCommandColumn" HeaderText="Edit" ItemStyle-Width="10%"> </telerik:GridEditCommandColumn> <telerik:GridButtonColumn CommandName="Delete" Text="Delete" UniqueName="DeleteColumn" ConfirmText="Are you sure you want to delete this application?" ConfirmTitle="Confirm Delete" ConfirmDialogType="Classic" ItemStyle-Width="10%" HeaderText="Delete"> </telerik:GridButtonColumn> <telerik:GridBoundColumn DataField="AppId" UniqueName="AppId" Visible="false" HeaderText="Application Id" ReadOnly="true"> </telerik:GridBoundColumn> <telerik:GridBoundColumn DataField="AppName" UniqueName="AppName" HeaderText="Application Name" MaxLength="30" ItemStyle-Width="40%"> </telerik:GridBoundColumn> <telerik:GridTemplateColumn DataField="ServerId" UniqueName="ServerId" HeaderText="Server Hosted" EditFormColumnIndex="1"> <EditItemTemplate> <asp:DropDownList ID="ddlServerHosted" runat="server" DataTextField="Servers.ServerName" DataValueField="ServerId" Width="40%"> </asp:DropDownList> </EditItemTemplate> </telerik:GridTemplateColumn> </Columns> <EditFormSettings ColumnNumber="2" CaptionDataField="AppId" InsertCaption="Insert New Application" EditFormType="AutoGenerated"> <EditColumn InsertText="Insert record" EditText="Edit application id #:" EditFormColumnIndex="0" UpdateText="Application updated" UniqueName="InsertCommandColumn1" CancelText="Cancel insert" ButtonType="ImageButton"></EditColumn> <FormTableItemStyle Wrap="false" /> <FormTableStyle GridLines="Horizontal" CellPadding="2" CellSpacing="0" Height="110px" Width="110px" /> <FormTableAlternatingItemStyle Wrap="false" /> <FormStyle Width="100%" BackColor="#EEF2EA" /> <FormTableButtonRowStyle HorizontalAlign="Right" /> </EditFormSettings> </MasterTableView> </telerik:RadGrid> <asp:EntityDataSource ID="applicationsEntityDataSource" runat="server" ConnectionString="name=AnalyticsEntities" EnableDelete="True" EntityTypeFilter="Applications" EnableInsert="True" EnableUpdate="True" EntitySetName="Applications" DefaultContainerName="AnalyticsEntities" Include="Servers"> </asp:EntityDataSource> I tried another approach where I replaced the GridTemplateColumn with the following code <telerik:RadComboBox ID="RadComboBox1" DataSourceID="serversEntityDataSource" DataTextField="ServerName" DataValueField="ServerId" AppendDataBoundItems="true" runat="server" > <Items> <telerik:RadComboBoxItem /> </Items> and using a separate EntityDataSource control as follows: <asp:EntityDataSource ID="serversEntityDataSource" runat="server" ConnectionString="name=AnalyticsEntities" EnableDelete="True" EntityTypeFilter="Servers" EnableInsert="True" EnableUpdate="True" EntitySetName="Servers" DefaultContainerName="AnalyticsEntities"> </asp:EntityDataSource> but, I get the following error. Application cannot be inserted. Reason: Entities in 'AnalyticsEntities.Applications' participate in the 'FK_Servers_Applications' relationship. 0 related 'Servers' were found. 1 'Servers' is expected. My question is, how do you bind the navigation property and load the values in the DropDownList/RadComboBox control?

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  • dynatree: how can i select child node programmatically

    - by Muhammad Adeel Zahid
    hello everyone i m using jquery's dynaTree in my application and i want to select the all the child nodes programmably when a node is selected. the structure of my tree is as follows <div id = "tree"> <ul> <li>package 1 <ul> <li>module 1.1 <ul> <li> document 1.1.1</li> <li> document 1.1.2</li> </ul> </li> <li>module 1.2 <ul> <li>document 1.2.1</li> <li>document 1.2.2</li> </ul> </li> </ul> </li> <li> package 2 <ul> <li> module 2.1 <ul> <li>document 2.1.1</li> <li>document 2.1.1</li> </ul> </li> </ul> </li> </ul> </div> now what i want is that when i click on tree node with title "package 1" all its child nodes i.e (module 1.1, document 1.1.1, document 1.1.2, module 1.2, document 1.2.1, document 1.2.2) should also be selected below is the approach i tried to use $("#tree").dynatree({ onSelect: function(flag, dtnode) { // This will happen each time a check box is selected/deselected var selectedNodes = dtnode.tree.getSelectedNodes(); var selectedKeys = $.map(selectedNodes, function(node) { //alert(node.data.key); return node.data.key; }); // Set the hidden input field's value to the selected items $('#SelectedItems').val(selectedKeys.join(",")); if (flag) { child = dtnode.childList; alert(child.length); for (i = 0; i < child.length; i++) { var x = child[i].select(true); alert(i); } } }, checkbox: true, onActivate: function(dtnode) { //alert("You activated " + dtnode.data.key); } }); in the if(flag) condition i get all the child nodes of element that is selected by user and it gives me the correct value that i can see from alert(child.length) statement. then i run the loop to select all the children but loop never goes beyond the statement var x = child[i].select(true); and i can never see the statement alert(i) being executed. the result of above statement is that if i select package 1, module 1.1 and document 1.1.1 is also selected but never does it execute alert(i) statement neither other children of package 1 are selected. in my view when first time child[i].select(true) statement is executed it also triggers the on select event of its children thus making a recursion kind of thing is my thinking correct? no matter recursion or what why on earth does it not complete the loop and execute very next instruction alert(i). please help me in solving this problem. i m dying to see that alert any suggestion and help is highly appriciated thanks Adeel

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  • FluentNHibernate Unit Of Work / Repository Design Pattern Questions

    - by Echiban
    Hi all, I think I am at a impasse here. I have an application I built from scratch using FluentNHibernate (ORM) / SQLite (file db). I have decided to implement the Unit of Work and Repository Design pattern. I am at a point where I need to think about the end game, which will start as a WPF windows app (using MVVM) and eventually implement web services / ASP.Net as UI. Now I already created domain objects (entities) for ORM. And now I don't know how should I use it outside of ORM. Questions about it include: Should I use ORM entity objects directly as models in MVVM? If yes, do I put business logic (such as certain values must be positive and be greater than another Property) in those entity objects? It is certainly the simpler approach, and one I am leaning right now. However, will there be gotchas that would trash this plan? If the answer above is no, do I then create a new set of classes to implement business logic and use those as Models in MVVM? How would I deal with the transition between model objects and entity objects? I guess a type converter implementation would work well here. Now I followed this well written article to implement the Unit Of Work pattern. However, due to the fact that I am using FluentNHibernate instead of NHibernate, I had to bastardize the implementation of UnitOfWorkFactory. Here's my implementation: using System; using FluentNHibernate.Cfg; using FluentNHibernate.Cfg.Db; using NHibernate; using NHibernate.Cfg; using NHibernate.Tool.hbm2ddl; namespace ELau.BlindsManagement.Business { public class UnitOfWorkFactory : IUnitOfWorkFactory { private static readonly string DbFilename; private static Configuration _configuration; private static ISession _currentSession; private ISessionFactory _sessionFactory; static UnitOfWorkFactory() { // arbitrary default filename DbFilename = "defaultBlindsDb.db3"; } internal UnitOfWorkFactory() { } #region IUnitOfWorkFactory Members public ISession CurrentSession { get { if (_currentSession == null) { throw new InvalidOperationException(ExceptionStringTable.Generic_NotInUnitOfWork); } return _currentSession; } set { _currentSession = value; } } public ISessionFactory SessionFactory { get { if (_sessionFactory == null) { _sessionFactory = BuildSessionFactory(); } return _sessionFactory; } } public Configuration Configuration { get { if (_configuration == null) { Fluently.Configure().ExposeConfiguration(c => _configuration = c); } return _configuration; } } public IUnitOfWork Create() { ISession session = CreateSession(); session.FlushMode = FlushMode.Commit; _currentSession = session; return new UnitOfWorkImplementor(this, session); } public void DisposeUnitOfWork(UnitOfWorkImplementor adapter) { CurrentSession = null; UnitOfWork.DisposeUnitOfWork(adapter); } #endregion public ISession CreateSession() { return SessionFactory.OpenSession(); } public IStatelessSession CreateStatelessSession() { return SessionFactory.OpenStatelessSession(); } private static ISessionFactory BuildSessionFactory() { ISessionFactory result = Fluently.Configure() .Database( SQLiteConfiguration.Standard .UsingFile(DbFilename) ) .Mappings(m => m.FluentMappings.AddFromAssemblyOf<UnitOfWorkFactory>()) .ExposeConfiguration(BuildSchema) .BuildSessionFactory(); return result; } private static void BuildSchema(Configuration config) { // this NHibernate tool takes a configuration (with mapping info in) // and exports a database schema from it _configuration = config; new SchemaExport(_configuration).Create(false, true); } } } I know that this implementation is flawed because a few tests pass when run individually, but when all tests are run, it would fail for some unknown reason. Whoever wants to help me out with this one, given its complexity, please contact me by private message. I am willing to send some $$$ by Paypal to someone who can address the issue and provide solid explanation. I am new to ORM, so any assistance is appreciated.

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  • Paying great programmers more than average programmers

    - by Kelly French
    It's fairly well recognized that some programmers are up to 10 times more productive than others. Joel mentions this topic on his blog. There is a whole blog devoted to the idea of the "10x productive programmer". In years since the original study, the general finding that "There are order-of-magnitude differences among programmers" has been confirmed by many other studies of professional programmers (Curtis 1981, Mills 1983, DeMarco and Lister 1985, Curtis et al. 1986, Card 1987, Boehm and Papaccio 1988, Valett and McGarry 1989, Boehm et al 2000). Fred Brooks mentions the wide range in the quality of designers in his "No Silver Bullet" article, The differences are not minor--they are rather like the differences between Salieri and Mozart. Study after study shows that the very best designers produce structures that are faster, smaller, simpler, cleaner, and produced with less effort. The differences between the great and the average approach an order of magnitude. The study that Brooks cites is: H. Sackman, W.J. Erikson, and E.E. Grant, "Exploratory Experimental Studies Comparing Online and Offline Programming Performance," Communications of the ACM, Vol. 11, No. 1 (January 1968), pp. 3-11. The way programmers are paid by employers these days makes it almost impossible to pay the great programmers a large multiple of what the entry-level salary is. When the starting salary for a just-graduated entry-level programmer, we'll call him Asok (From Dilbert), is $40K, even if the top programmer, we'll call him Linus, makes $120K that is only a multiple of 3. I'd be willing to be that Linus does much more than 3 times what Asok does, so why wouldn't we expect him to get paid more as well? Here is a quote from Stroustrup: "The companies are complaining because they are hurting. They can't produce quality products as cheaply, as reliably, and as quickly as they would like. They correctly see a shortage of good developers as a part of the problem. What they generally don't see is that inserting a good developer into a culture designed to constrain semi-skilled programmers from doing harm is pointless because the rules/culture will constrain the new developer from doing anything significantly new and better." This leads to two questions. I'm excluding self-employed programmers and contractors. If you disagree that's fine but please include your rationale. It might be that the self-employed or contract programmers are where you find the top-10 earners, but please provide a explanation/story/rationale along with any anecdotes. [EDIT] I thought up some other areas in which talent/ability affects pay. Financial traders (commodities, stock, derivatives, etc.) designers (fashion, interior decorators, architects, etc.) professionals (doctor, lawyer, accountant, etc.) sales Questions: Why aren't the top 1% of programmers paid like A-list movie stars? What would the industry be like if we did pay the "Smart and gets things done" programmers 6, 8, or 10 times what an intern makes? [Footnote: I posted this question after submitting it to the Stackoverflow podcast. It was included in episode 77 and I've written more about it as a Codewright's Tale post 'Of Rockstars and Bricklayers'] Epilogue: It's probably unfair to exclude contractors and the self-employed. One aspect of the highest earners in other fields is that they are free-agents. The competition for their skills is what drives up their earning power. This means they can not be interchangeable or otherwise treated as a plug-and-play resource. I liked the example in one answer of a major league baseball team trying to field two first-basemen. Also, something that Joel mentioned in the Stackoverflow podcast (#77). There are natural dynamics to shrink any extreme performance/pay ranges between the highs and lows. One is the peer pressure of organizations to pay within a given range, another is the likelyhood that the high performer will realize their undercompensation and seek greener pastures.

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  • Combined SOAP/JSON/XML in WCF, using UriTemplate

    - by gregmac
    I'm trying to build a generic web service interface using WCF, to allow 3rd party developers to hook into our software. After much struggling and reading (this question helped a lot), I finally got SOAP, JSON and XML (POX) working together. To simplify, here's my code (to make this example simple, I'm not using interfaces -- I did try this both ways): <ServiceContract()> _ Public Class TestService Public Sub New() End Sub <OperationContract()> _ <WebGet()> _ Public Function GetDate() As DateTime Return Now End Function '<WebGet(UriTemplate:="getdateoffset/{numDays}")> _ <OperationContract()> _ Public Function GetDateOffset(ByVal numDays As Integer) As DateTime Return Now.AddDays(numDays) End Function End Class and the web.config code: <services> <service name="TestService" behaviorConfiguration="TestServiceBehavior"> <endpoint address="soap" binding="basicHttpBinding" contract="TestService"/> <endpoint address="json" binding="webHttpBinding" behaviorConfiguration="jsonBehavior" contract="TestService"/> <endpoint address="xml" binding="webHttpBinding" behaviorConfiguration="poxBehavior" contract="TestService"/> <endpoint address="mex" contract="IMetadataExchange" binding="mexHttpBinding" /> </service> </services> <behaviors> <endpointBehaviors> <behavior name="jsonBehavior"> <enableWebScript/> </behavior> <behavior name="poxBehavior"> <webHttp /> </behavior> </endpointBehaviors> <serviceBehaviors> <behavior name="TestServiceBehavior"> <serviceMetadata httpGetEnabled="true"/> <serviceDebug includeExceptionDetailInFaults="false"/> </behavior> </serviceBehaviors> </behaviors> This actually works -- I'm able to go to TestService.svc/xml/GetDate for xml, TestService.svc/json/GetDate for json, and point a SOAP client at TestService.svc?wsdl and have the SOAP queries work. The part I'd like to fix is the queries. I have to use TestService.svc/xml/GetDateOffset?numDays=4 instead of TestService.svc/xml/GetDateOffset/4. If I specify the UriTemplate, I get the error: Endpoints using 'UriTemplate' cannot be used with 'System.ServiceModel.Description.WebScriptEnablingBehavior'. But of course without using <enableWebScript/>, JSON doesn't work. The only other thing I've seen that I think will work is making 3 different services (.svc files), that all implement an interface that specifies the contract, but in the classes specify different WebGet/WebInvoke attributes on each class. This seems like a lot of extra work, that frankly, I don't see why the framework doesn't handle for me. The implementation of the classes would all be the same, except for the attributes, which means over time it would be easy for bugs/changes to get fixed/done in one implementation but not the others, leading to inconsistent behaviour when using the JSON vs SOAP implementation for example. Am I doing something wrong here? Am I taking a totally wrong approach and misusing WCF? Is there a better way to do this? With my experience doing web stuff, I think it should be possible for some kind of framework to handle this ... I even have an idea in my head of how to build it. It just seems like WCF is supposed to be doing this, and I don't really want to reinvent the wheel.

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  • Edit and render RichText

    - by OregonGhost
    We have an application (a custom network management tool for building automation) that supports printing labels that you can cut out and insert into the devices' front displays. In earlier versions of the tool (not developed in my company), the application just pushed the strings into an Excel file that the field technician could then manipulate (like formatting text). We didn't do this in the new version because it was hard (impossible) to keep the Excel file in sync, and to avoid a binding to an external application (let alone different versions of Excel). We're using PDFSharp for rendering the labels. It has a System.Drawing-like interface, but can output to a System.Drawing.Graphics (screen / printer) as well as to a PDF file, which is a requirement. Later, basic formatting was introduced like Font Family, Style, Size, Color which would apply to one label (i.e. to exactly one string). Now the customer wants to be able to apply these formats to single characters in a string. I think the easiest way would be to support a subset of RichText. It's not as easy as I thought though. Currently the editor just displays a TextBox for the label you want to edit, with the font set to the label's font. I thought I'd just replace it with RichTextBox, and update the formatting buttons to use the RichTextBox formatting properties. Fairly easy. However, I need to draw the text. I know you can get the RichTextBox to draw to a HDC or System.Drawing.Graphics - but as already said, I need it to use PDFSharp. Rendering to bitmaps is not an option, since the PDF must not be huge, and it's a lot of labels. Unfortunately I couldn't get the RichTextBox to tell me the layout of the text - I'm fine with doing the actual rendering by hand, as long as I know where to draw what. This is the first question: How can I get the properly layouted metrics of the rich text out of a RichTextBox? Or is there any way to convert the rich text to a vector graphics format that can be easily drawn manually? I know about NRTFTree which can be used to parse and manipulate RichText. The documentation is bad (actually I don't know, it's Spanish), but I think I can get it to work. As far as I understood, it won't provide layouting as well. Because of this, I think I'll have to write a custom edit control (remember, it's basically just one or two line labels with basic RTF formatting, not a full-fledged edit control - more like editing a textbox in PowerPoint) and write custom text layout logic that used PDFSharp rather than System.Drawing for drawing. Is there any existing, even if partial, solution available, either for the editing or for doing the layout manually (or both)? Or is there an entirely different approach I'm just not seeing? Bonus points if exporting the label texts as RTF into a CSV file, and then importing in Excel retains the formatting. For the editing part, I need it to work in Windows Forms. Other than that it's not Windows-Forms-related, I think.

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  • Android: OutOfMemoryError while uploading video...

    - by AP257
    Hi all, I have the same problem as described here, but I will supply a few more details. While trying to upload a video in Android, I'm reading it into memory, and if the video is large I get an OutOfMemoryError. Here's my code: // get bytestream to upload videoByteArray = getBytesFromFile(cR, fileUriString); public static byte[] getBytesFromFile(ContentResolver cR, String fileUriString) throws IOException { Uri tempuri = Uri.parse(fileUriString); InputStream is = cR.openInputStream(tempuri); byte[] b3 = readBytes(is); is.close(); return b3; } public static byte[] readBytes(InputStream inputStream) throws IOException { ByteArrayOutputStream byteBuffer = new ByteArrayOutputStream(); // this is storage overwritten on each iteration with bytes int bufferSize = 1024; byte[] buffer = new byte[bufferSize]; int len = 0; while ((len = inputStream.read(buffer)) != -1) { byteBuffer.write(buffer, 0, len); } return byteBuffer.toByteArray(); } And here's the traceback (the error is thrown on the byteBuffer.write(buffer, 0, len) line): 04-08 11:56:20.456: ERROR/dalvikvm-heap(6088): Out of memory on a 16775184-byte allocation. 04-08 11:56:20.456: INFO/dalvikvm(6088): "IntentService[UploadService]" prio=5 tid=17 RUNNABLE 04-08 11:56:20.456: INFO/dalvikvm(6088): | group="main" sCount=0 dsCount=0 s=N obj=0x449a3cf0 self=0x38d410 04-08 11:56:20.456: INFO/dalvikvm(6088): | sysTid=6119 nice=0 sched=0/0 cgrp=default handle=4010416 04-08 11:56:20.456: INFO/dalvikvm(6088): at java.io.ByteArrayOutputStream.expand(ByteArrayOutputStream.java:~93) 04-08 11:56:20.456: INFO/dalvikvm(6088): at java.io.ByteArrayOutputStream.write(ByteArrayOutputStream.java:218) 04-08 11:56:20.456: INFO/dalvikvm(6088): at com.android.election2010.UploadService.readBytes(UploadService.java:199) 04-08 11:56:20.456: INFO/dalvikvm(6088): at com.android.election2010.UploadService.getBytesFromFile(UploadService.java:182) 04-08 11:56:20.456: INFO/dalvikvm(6088): at com.android.election2010.UploadService.doUploadinBackground(UploadService.java:118) 04-08 11:56:20.456: INFO/dalvikvm(6088): at com.android.election2010.UploadService.onHandleIntent(UploadService.java:85) 04-08 11:56:20.456: INFO/dalvikvm(6088): at android.app.IntentService$ServiceHandler.handleMessage(IntentService.java:30) 04-08 11:56:20.456: INFO/dalvikvm(6088): at android.os.Handler.dispatchMessage(Handler.java:99) 04-08 11:56:20.456: INFO/dalvikvm(6088): at android.os.Looper.loop(Looper.java:123) 04-08 11:56:20.456: INFO/dalvikvm(6088): at android.os.HandlerThread.run(HandlerThread.java:60) 04-08 11:56:20.467: WARN/dalvikvm(6088): threadid=17: thread exiting with uncaught exception (group=0x4001b180) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): Uncaught handler: thread IntentService[UploadService] exiting due to uncaught exception 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): java.lang.OutOfMemoryError 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at java.io.ByteArrayOutputStream.expand(ByteArrayOutputStream.java:93) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at java.io.ByteArrayOutputStream.write(ByteArrayOutputStream.java:218) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at com.android.election2010.UploadService.readBytes(UploadService.java:199) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at com.android.election2010.UploadService.getBytesFromFile(UploadService.java:182) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at com.android.election2010.UploadService.doUploadinBackground(UploadService.java:118) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at com.android.election2010.UploadService.onHandleIntent(UploadService.java:85) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at android.app.IntentService$ServiceHandler.handleMessage(IntentService.java:30) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at android.os.Handler.dispatchMessage(Handler.java:99) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at android.os.Looper.loop(Looper.java:123) 04-08 11:56:20.467: ERROR/AndroidRuntime(6088): at android.os.HandlerThread.run(HandlerThread.java:60) 04-08 11:56:20.496: INFO/Process(4657): Sending signal. PID: 6088 SIG: 3 I guess that as @DroidIn suggests, I need to upload it in chunks. But (newbie question alert) does that mean that I should make multiple PostMethod requests, and glue the file together at the server end? Or can I load the bytestream into memory in chunks, and glue it together in the Android code? If anyone could give me a clue as to the best approach, I would be very grateful.

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  • Point in polygon OR point on polygon using LINQ

    - by wageoghe
    As noted in an earlier question, How to Zip enumerable with itself, I am working on some math algorithms based on lists of points. I am currently working on point in polygon. I have the code for how to do that and have found several good references here on SO, such as this link Hit test. So, I can figure out whether or not a point is in a polygon. As part of determining that, I want to determine if the point is actually on the polygon. This I can also do. If I can do all of that, what is my question you might ask? Can I do it efficiently using LINQ? I can already do something like the following (assuming a Pairwise extension method as described in my earlier question as well as in links to which my question/answers links, and assuming a Position type that has X and Y members). I have not tested much, so the lambda might not be 100% correct. Also, it does not take very small differences into account. public static PointInPolygonLocation PointInPolygon(IEnumerable<Position> pts, Position pt) { int numIntersections = pts.Pairwise( (p1, p2) => { if (p1.Y != p2.Y) { if ((p1.Y >= pt.Y && p2.Y < pt.Y) || (p1.Y < pt.Y && p2.Y >= pt.Y)) { if (p1.X < p1.X && p2.X < pt.X) { return 1; } if (p1.X < pt.X || p2.X < pt.X) { if (((pt.Y - p1.Y) * ((p1.X - p2.X) / (p1.Y - p2.Y)) * p1.X) < pt.X) { return 1; } } } } return 0; }).Sum(); if (numIntersections % 2 == 0) { return PointInPolygonLocation.Outside; } else { return PointInPolygonLocation.Inside; } } This function, PointInPolygon, takes the input Position, pt, iterates over the input sequence of position values, and uses the Jordan Curve method to determine how many times a ray extended from pt to the left intersects the polygon. The lambda expression will yield, into the "zipped" list, 1 for every segment that is crossed, and 0 for the rest. The sum of these values determines if pt is inside or outside of the polygon (odd == inside, even == outside). So far, so good. Now, for any consecutive pairs of position values in the sequence (i.e. in any execution of the lambda), we can also determine if pt is ON the segment p1, p2. If that is the case, we can stop the calculation because we have our answer. Ultimately, my question is this: Can I perform this calculation (maybe using Aggregate?) such that we will only iterate over the sequence no more than 1 time AND can we stop the iteration if we encounter a segment that pt is ON? In other words, if pt is ON the very first segment, there is no need to examine the rest of the segments because we have the answer. It might very well be that this operation (particularly the requirement/desire to possibly stop the iteration early) does not really lend itself well to the LINQ approach. It just occurred to me that maybe the lambda expression could yield a tuple, the intersection value (1 or 0 or maybe true or false) and the "on" value (true or false). Maybe then I could use TakeWhile(anontype.PointOnPolygon == false). If I Sum the tuples and if ON == 1, then the point is ON the polygon. Otherwise, the oddness or evenness of the sum of the other part of the tuple tells if the point is inside or outside.

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  • Need help with setting up comet code

    - by Saif Bechan
    Does anyone know off a way or maybe think its possible to connect Node.js with Nginx http push module to maintain a persistent connection between client and browser. I am new to comet so just don't understand the publishing etc maybe someone can help me with this. What i have set up so far is the following. I downloaded the jQuery.comet plugin and set up the following basic code: Client JavaScript <script type="text/javascript"> function updateFeed(data) { $('#time').text(data); } function catchAll(data, type) { console.log(data); console.log(type); } $.comet.connect('/broadcast/sub?channel=getIt'); $.comet.bind(updateFeed, 'feed'); $.comet.bind(catchAll); $('#kill-button').click(function() { $.comet.unbind(updateFeed, 'feed'); }); </script> What I can understand from this is that the client will keep on listening to the url followed by /broadcast/sub=getIt. When there is a message it will fire updateFeed. Pretty basic and understandable IMO. Nginx http push module config default_type application/octet-stream; sendfile on; keepalive_timeout 65; push_authorized_channels_only off; server { listen 80; location /broadcast { location = /broadcast/sub { set $push_channel_id $arg_channel; push_subscriber; push_subscriber_concurrency broadcast; push_channel_group broadcast; } location = /broadcast/pub { set $push_channel_id $arg_channel; push_publisher; push_min_message_buffer_length 5; push_max_message_buffer_length 20; push_message_timeout 5s; push_channel_group broadcast; } } } Ok now this tells nginx to listen at port 80 for any calls to /broadcast/sub and it will give back any responses sent to /broadcast/pub. Pretty basic also. This part is not so hard to understand, and is well documented over the internet. Most of the time there is a ruby or a php file behind this that does the broadcasting. My idea is to have node.js broadcasting /broadcast/pub. I think this will let me have persistent streaming data from the server to the client without breaking the connection. I tried the long-polling approach with looping the request but I think this will be more efficient. Or is this not going to work. Node.js file Now to create the Node.js i'm lost. First off all I don't know how to have node.js to work in this way. The setup I used for long polling is as follows: var sys = require('sys'), http = require('http'); http.createServer(function (req, res) { res.writeHead(200, {'Content-Type': 'text/html'}); res.write(new Date()); res.close(); seTimeout('',1000); }).listen(8000); This listens to port 8000 and just writes on the response variable. For long polling my nginx.config looked something like this: server { listen 80; server_name _; location / { proxy_pass http://mydomain.com:8080$request_uri; include /etc/nginx/proxy.conf; } } This just redirected the port 80 to 8000 and this worked fine. Does anyone have an idea on how to have Node.js act in a way Comet understands it. Would be really nice and you will help me out a lot. Recources used An example where this is done with ruby instead of Node.js jQuery.comet Nginx HTTP push module homepage Faye: a Comet client and server for Node.js and Rack To use faye I have to install the comet client, but I want to use the one supplied with Nginx. Thats why I don't just use faye. The one nginx uses is much more optimzed. extra Persistant connections Going evented with Node.js

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  • Marrying Core Animation with OpenGL ES

    - by Ole Begemann
    Edit: I suppose instead of the long explanation below I might also ask: Sending -setNeedsDisplay to an instance of CAEAGLLayer does not cause the layer to redraw (i.e., -drawInContext: is not called). Instead, I get this console message: <GLLayer: 0x4d500b0>: calling -display has no effect. Is there a way around this issue? Can I invoke -drawInContext: when -setNeedsDisplay is called? Long explanation below: I have an OpenGL scene that I would like to animate using Core Animation animations. Following the standard approach to animate custom properties in a CALayer, I created a subclass of CAEAGLLayer and defined a property sceneCenterPoint in it whose value should be animated. My layer also holds a reference to the OpenGL renderer: #import <UIKit/UIKit.h> #import <QuartzCore/QuartzCore.h> #import "ES2Renderer.h" @interface GLLayer : CAEAGLLayer { ES2Renderer *renderer; } @property (nonatomic, retain) ES2Renderer *renderer; @property (nonatomic, assign) CGPoint sceneCenterPoint; I then declare the property @dynamic to let CA create the accessors, override +needsDisplayForKey: and implement -drawInContext: to pass the current value of the sceneCenterPoint property to the renderer and ask it to render the scene: #import "GLLayer.h" @implementation GLLayer @synthesize renderer; @dynamic sceneCenterPoint; + (BOOL) needsDisplayForKey:(NSString *)key { if ([key isEqualToString:@"sceneCenterPoint"]) { return YES; } else { return [super needsDisplayForKey:key]; } } - (void) drawInContext:(CGContextRef)ctx { self.renderer.centerPoint = self.sceneCenterPoint; [self.renderer render]; } ... (If you have access to the WWDC 2009 session videos, you can review this technique in session 303 ("Animated Drawing")). Now, when I create an explicit animation for the layer on the keyPath @"sceneCenterPoint", Core Animation should calculate the interpolated values for the custom properties and call -drawInContext: for each step of the animation: - (IBAction)animateButtonTapped:(id)sender { CABasicAnimation *animation = [CABasicAnimation animationWithKeyPath:@"sceneCenterPoint"]; animation.duration = 1.0; animation.fromValue = [NSValue valueWithCGPoint:CGPointZero]; animation.toValue = [NSValue valueWithCGPoint:CGPointMake(1.0f, 1.0f)]; [self.glView.layer addAnimation:animation forKey:nil]; } At least that is what would happen for a normal CALayer subclass. When I subclass CAEAGLLayer, I get this output on the console for each step of the animation: 2010-12-21 13:59:22.180 CoreAnimationOpenGL[7496:207] <GLLayer: 0x4e0be20>: calling -display has no effect. 2010-12-21 13:59:22.198 CoreAnimationOpenGL[7496:207] <GLLayer: 0x4e0be20>: calling -display has no effect. 2010-12-21 13:59:22.216 CoreAnimationOpenGL[7496:207] <GLLayer: 0x4e0be20>: calling -display has no effect. 2010-12-21 13:59:22.233 CoreAnimationOpenGL[7496:207] <GLLayer: 0x4e0be20>: calling -display has no effect. ... So it seems that, possibly for performance reasons, for OpenGL layers, -drawInContext: is not getting called because these layers do not use the standard -display method to draw themselves. Can anybody confirm that? Is there a way around it? Or can I not use the technique I laid out above? This would mean I would have to implement the animations manually in the OpenGL renderer (which is possible but not as elegant IMO).

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