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  • CLR 4.0 inlining policy? (maybe bug with MethodImplOptions.NoInlining)

    - by ControlFlow
    I've testing some new CLR 4.0 behavior in method inlining (cross-assembly inlining) and found some strage results: Assembly ClassLib.dll: using System.Diagnostics; using System; using System.Reflection; using System.Security; using System.Runtime.CompilerServices; namespace ClassLib { public static class A { static readonly MethodInfo GetExecuting = typeof(Assembly).GetMethod("GetExecutingAssembly"); public static Assembly Foo(out StackTrace stack) // 13 bytes { // explicit call to GetExecutingAssembly() stack = new StackTrace(); return Assembly.GetExecutingAssembly(); } public static Assembly Bar(out StackTrace stack) // 25 bytes { // reflection call to GetExecutingAssembly() stack = new StackTrace(); return (Assembly) GetExecuting.Invoke(null, null); } public static Assembly Baz(out StackTrace stack) // 9 bytes { stack = new StackTrace(); return null; } public static Assembly Bob(out StackTrace stack) // 13 bytes { // call of non-inlinable method! return SomeSecurityCriticalMethod(out stack); } [SecurityCritical, MethodImpl(MethodImplOptions.NoInlining)] static Assembly SomeSecurityCriticalMethod(out StackTrace stack) { stack = new StackTrace(); return Assembly.GetExecutingAssembly(); } } } Assembly ConsoleApp.exe using System; using ClassLib; using System.Diagnostics; class Program { static void Main() { Console.WriteLine("runtime: {0}", Environment.Version); StackTrace stack; Console.WriteLine("Foo: {0}\n{1}", A.Foo(out stack), stack); Console.WriteLine("Bar: {0}\n{1}", A.Bar(out stack), stack); Console.WriteLine("Baz: {0}\n{1}", A.Baz(out stack), stack); Console.WriteLine("Bob: {0}\n{1}", A.Bob(out stack), stack); } } Results: runtime: 4.0.30128.1 Foo: ClassLib, Version=1.0.0.0, Culture=neutral, PublicKeyToken=null at ClassLib.A.Foo(StackTrace& stack) at Program.Main() Bar: ClassLib, Version=1.0.0.0, Culture=neutral, PublicKeyToken=null at ClassLib.A.Bar(StackTrace& stack) at Program.Main() Baz: at Program.Main() Bob: ClassLib, Version=1.0.0.0, Culture=neutral, PublicKeyToken=null at Program.Main() So questions are: Why JIT does not inlined Foo and Bar calls as Baz does? They are lower than 32 bytes of IL and are good candidates for inlining. Why JIT inlined call of Bob and inner call of SomeSecurityCriticalMethod that is marked with the [MethodImpl(MethodImplOptions.NoInlining)] attribute? Why GetExecutingAssembly returns a valid assembly when is called by inlined Baz and SomeSecurityCriticalMethod methods? I've expect that it performs the stack walk to detect the executing assembly, but stack will contains only Program.Main() call and no methods of ClassLib assenbly, to ConsoleApp should be returned.

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  • How to combine designable components with dependency injection

    - by Wim Coenen
    When creating a designable .NET component, you are required to provide a default constructor. From the IComponent documentation: To be a component, a class must implement the IComponent interface and provide a basic constructor that requires no parameters or a single parameter of type IContainer. This makes it impossible to do dependency injection via constructor arguments. (Extra constructors could be provided, but the designer would ignore them.) Some alternatives we're considering: Service Locator Don't use dependency injection, instead use the service locator pattern to acquire dependencies. This seems to be what IComponent.Site.GetService is for. I guess we could create a reusable ISite implementation (ConfigurableServiceLocator?) which can be configured with the necessary dependencies. But how does this work in a designer context? Dependency Injection via properties Inject dependencies via properties. Provide default instances if they are necessary to show the component in a designer. Document which properties need to be injected. Inject dependencies with an Initialize method This is much like injection via properties but it keeps the list of dependencies that need to be injected in one place. This way the list of required dependencies is documented implicitly, and the compiler will assists you with errors when the list changes. Any idea what the best practice is here? How do you do it? edit: I have removed "(e.g. a WinForms UserControl)" since I intended the question to be about components in general. Components are all about inversion of control (see section 8.3.1 of the UMLv2 specification) so I don't think that "you shouldn't inject any services" is a good answer. edit 2: It took some playing with WPF and the MVVM pattern to finally "get" Mark's answer. I see now that visual controls are indeed a special case. As for using non-visual components on designer surfaces, I think the .NET component model is fundamentally incompatible with dependency injection. It appears to be designed around the service locator pattern instead. Maybe this will start to change with the infrastructure that was added in .NET 4.0 in the System.ComponentModel.Composition namespace.

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  • How do access a secure website within a sharepoint webpart?

    - by Bill
    How do access a secure website within a sharepoint webpart? The following code works fine as a console application but if you run it in a webpart, you will get a access violation WebRequest request = WebRequest.Create("https://somesecuresite.com"); WebResponse firstResponse = null; try { firstResponse = request.GetResponse(); } catch (WebException ex) { writer.WriteLine("Error: " + ex.ToString()); return; } if you access a non secure site, it also works. Any ideas? Error: System.Net.WebException: The underlying connection was closed: An unexpected error occurred on a receive. --- System.AccessViolationException: Attempted to read or write protected memory. This is often an indication that other memory is corrupt. at System.Net.UnsafeNclNativeMethods.NativePKI.CertVerifyCertificateChainPolicy(IntPtr policy, SafeFreeCertChain chainContext, ChainPolicyParameter& cpp, ChainPolicyStatus& ps) at System.Net.PolicyWrapper.VerifyChainPolicy(SafeFreeCertChain chainContext, ChainPolicyParameter& cpp) at System.Net.Security.SecureChannel.VerifyRemoteCertificate(RemoteCertValidationCallback remoteCertValidationCallback) at System.Net.Security.SslState.CompleteHandshake() at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessReceivedBlob(Byte[] buffer, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReadFrame(Byte[] buffer, Int32 readBytes, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartReceiveBlob(Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.CheckCompletionBeforeNextReceive(ProtocolToken message, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.StartSendBlob(Byte[] incoming, Int32 count, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ForceAuthentication(Boolean receiveFirst, Byte[] buffer, AsyncProtocolRequest asyncRequest) at System.Net.Security.SslState.ProcessAuthentication(LazyAsyncResult lazyResult) at System.Net.TlsStream.CallProcessAuthentication(Object state) at System.Threading.ExecutionContext.runTryCode(Object userData) at System.Runtime.CompilerServices.RuntimeHelpers.ExecuteCodeWithGuaranteedCleanup(TryCode code, CleanupCode backoutCode, Object userData) at System.Threading.ExecutionContext.RunInternal(ExecutionContext executionContext, ContextCallback callback, Object state) at System.Threading.ExecutionContext.Run(ExecutionContext executionContext, ContextCallback callback, Object state) at System.Net.TlsStream.ProcessAuthentication(LazyAsyncResult result) at System.Net.TlsStream.Write(Byte[] buffer, Int32 offset, Int32 size) at System.Net.PooledStream.Write(Byte[] buffer, Int32 offset, Int32 size) at System.Net.ConnectStream.WriteHeaders(Boolean async) --- End of inner exception stack trace --- at System.Net.HttpWebRequest.GetResponse()

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  • Excel UDF calculation should return 'original' value

    - by LeChe
    Hi all, I've been struggling with a VBA problem for a while now and I'll try to explain it as thoroughly as possible. I have created a VSTO plugin with my own RTD implementation that I am calling from my Excel sheets. To avoid having to use the full-fledged RTD syntax in the cells, I have created a UDF that hides that API from the sheet. The RTD server I created can be enabled and disabled through a button in a custom Ribbon component. The behavior I want to achieve is as follows: If the server is disabled and a reference to my function is entered in a cell, I want the cell to display Disabled If the server is disabled, but the function had been entered in a cell when it was enabled (and the cell thus displays a value), I want the cell to keep displaying that value If the server is enabled, I want the cell to display Loading Sounds easy enough. Here is an example of the - non functional - code: Public Function RetrieveData(id as Long) Dim result as String // This returns either 'Disabled' or 'Loading' result = Application.Worksheet.Function.RTD("SERVERNAME", "", id) RetrieveData = result If(result = "Disabled") Then // Obviously, this recurses (and fails), so that's not an option If(Not IsEmpty(Application.Caller.Value2)) Then // So does this RetrieveData = Application.Caller.Value2 End If End If End Function The function will be called in thousands of cells, so storing the 'original' values in another data structure would be a major overhead and I would like to avoid it. Also, the RTD server does not know the values, since it also does not keep a history of it, more or less for the same reason. I was thinking that there might be some way to exit the function which would force it to not change the displayed value, but so far I have been unable to find anything like that. Any ideas on how to solve this are greatly appreciated! Thanks, Che EDIT: By popular demand, some additional info on why I want to do all this: As I said, the function will be called in thousands of cells and the RTD server needs to retrieve quite a bit of information. This can be quite hard on both network and CPU. To allow the user to decide for himself whether he wants this load on his machine, he or she can disable the updates from the server. In that case, he or she should still be able to calculate the sheets with the values currently in the fields, yet no updates are pushed into them. Once new data is required, the server can be enabled and the fields will be updated. Again, since we are talking about quite a bit of data here, I would rather not store it somewhere in the sheet. Plus, the data should be usable even if the workbook is closed and loaded again.

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  • symfony/zend integration - blank screen

    - by user142176
    Hi, I need to use ZendAMF on a symfony project and I'm currently working on integrating the two. I have a frontend app with two modules, one of which is 'gateway' - the AMF gateway. In my frontend app config, I have the following in the configure function: // load symfony autoloading first parent::initialize(); // Integrate Zend Framework require_once('[MY PATH TO ZEND]\Loader.php'); spl_autoload_register(array('Zend_Loader', 'autoload')); The executeIndex function my the gateway actions.class.php looks like this // No Layout $this->setLayout(false); // Set MIME Type $this->getResponse()->setContentType('application/x-amf; charset='.sfConfig::get('sf_charset')); // Disable cause this is a non-html page sfConfig::set('sf_web_debug', false); // Create AMF Server $server = new Zend_Amf_Server(); $server->setClass('MYCLASS'); echo $server->handle(); return sfView::NONE; Now when I try to visit the url for the gateway module, or even the other module which was working perfectly fine until this attempt, I only see a blank screen, with not even the symfony dev bar loaded. Oddly enough, my symfony logs are not being updated as well, which suggests that Synfony is not even being 'reached'. So presumably the error has something to do with Zend, but I have no idea how to figure out what the error could be. One thing I do know for sure is that this is not a file path error, because if I change the path in the following line (a part of frontendConfiguration as shown above), I get a Zend_Amf_Server not found error. So the path must be correct. Also if I comment out this very same line, the second module resumes to normality, and my gateway broadcasts a blank x-amf stream. spl_autoload_register(array('Zend_Loader', 'autoload')); Does anyone have any tips on how I could attach this problem? Thanks P.S. I'm currently running an older version of Zend, which is why I am using Zend_Loader instead of Zend_autoLoader (I think). But I've tried switching to the new lib, but the error still remains. So it's not a version problem as well.

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  • Is READ UNCOMMITTED / NOLOCK safe in this situation?

    - by Ben Challenor
    I know that snapshot isolation would fix this problem, but I'm wondering if NOLOCK is safe in this specific case so that I can avoid the overhead. I have a table that looks something like this: drop table Data create table Data ( Id BIGINT NOT NULL, Date BIGINT NOT NULL, Value BIGINT, constraint Cx primary key (Date, Id) ) create nonclustered index Ix on Data (Id, Date) There are no updates to the table, ever. Deletes can occur but they should never contend with the SELECT because they affect the other, older end of the table. Inserts are regular and page splits to the (Id, Date) index are extremely common. I have a deadlock situation between a standard INSERT and a SELECT that looks like this: select top 1 Date, Value from Data where Id = @p0 order by Date desc because the INSERT acquires a lock on Cx (Date, Id; Value) and then Ix (Id, Date), but the SELECT acquires a lock on Ix (Id, Date) and then Cx (Date, Id; Value). This is because the SELECT first seeks on Ix and then joins to a seek on Cx. Swapping the clustered and non-clustered index would break this cycle, but it is not an acceptable solution because it would introduce cycles with other (more complex) SELECTs. If I add NOLOCK to the SELECT, can it go wrong in this case? Can it return: More than one row, even though I asked for TOP 1? No rows, even though one exists and has been committed? Worst of all, a row that doesn't satisfy the WHERE clause? I've done a lot of reading about this online, but the only reproductions of over- or under-count anomalies I've seen (one, two) involve a scan. This involves only seeks. Jeff Atwood has a post about using NOLOCK that generated a good discussion. I was particularly interested in a comment by Rick Townsend: Secondly, if you read dirty data, the risk you run is of reading the entirely wrong row. For example, if your select reads an index to find your row, then the update changes the location of the rows (e.g.: due to a page split or an update to the clustered index), when your select goes to read the actual data row, it's either no longer there, or a different row altogether! Is this possible with inserts only, and no updates? If so, then I guess even my seeks on an insert-only table could be dangerous. Update: I'm trying to figure out how snapshot isolation works. It seems to be row-based, where transactions read the table (with no shared lock!), find the row they are interested in, and then see if they need to get an old version of the row from the version store in tempdb. But in my case, no row will have more than one version, so the version store seems rather pointless. And if the row was found with no shared lock, how is it different to just using NOLOCK?

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  • Linq to sql C# updating reference Tables

    - by Laurence Burke
    ok reclarification I am adding a new address and I know the structure as AddressID = PK and all other entities are non nullable. Now on insert of a new row the addrID Pk is autogened and I am wondering if I would have to get that to create a new row in the referencing table or does that automatically get generated also. also I want to be able to repopulate the dropdownlist that lists the current employee's addresses with the newly created address. static uint _curEmpID; protected void btnAdd_Click(object sender, EventArgs e) { if (txtZip.Text != "" && txtAdd1.Text != "" && txtCity.Text != "") { TestDataClassDataContext dc = new TestDataClassDataContext(); Address addr = new Address() { AddressLine1 = txtAdd1.Text, AddressLine2 = txtAdd2.Text, City = txtCity.Text, PostalCode = txtZip.Text, StateProvinceID = Convert.ToInt32(ddlState.SelectedValue) }; dc.Addresses.InsertOnSubmit(addr); lblSuccess.Visible = true; lblErrMsg.Visible = false; dc.SubmitChanges(); // // TODO: add reference from new address to CurEmp Table // SetAddrList(); } else { lblErrMsg.Text = "Invalid Input"; lblErrMsg.Visible = true; } } protected void ddlAddList_SelectedIndexChanged(object sender, EventArgs e) { lblErrMsg.Visible = false; lblSuccess.Visible = false; TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; if (ddlAddList.SelectedValue != "-1") { var addr = (from a in dc.Addresses where a.AddressID == Convert.ToInt32(ddlAddList.SelectedValue) select a).FirstOrDefault(); txtAdd1.Text = addr.AddressLine1; txtAdd2.Text = addr.AddressLine2; txtCity.Text = addr.City; txtZip.Text = addr.PostalCode; ddlState.SelectedValue = addr.StateProvinceID.ToString(); btnSubmit.Visible = true; btnAdd.Visible = false; } else { txtAdd1.Text = ""; txtAdd2.Text = ""; txtCity.Text = ""; txtZip.Text = ""; btnAdd.Visible = true; btnSubmit.Visible = false; } } protected void SetAddrList() { TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; var addList = from addr in dc.Addresses from eaddr in dc.EmployeeAddresses where eaddr.EmployeeID == _curEmpID && addr.AddressID == eaddr.AddressID select new { AddValue = addr.AddressID, AddText = addr.AddressID, }; ddlAddList.DataSource = addList; ddlAddList.DataValueField = "AddValue"; ddlAddList.DataTextField = "AddText"; ddlAddList.DataBind(); ddlAddList.Items.Add(new ListItem("<Add Address>", "-1")); } OK I am hoping that I did not include too much code. I would really appreciate any other comments about I could otherwise improve this code in any other ways also.

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  • SQL efficiency argument, add a column or solvable by query?

    - by theTurk
    I am a recent college graduate and a new hire for software development. Things have been a little slow lately so I was given a db task. My db skills are limited to pet projects with Rails and Django. So, I was a little surprised with my latest task. I have been asked by my manager to subclass Person with a 'Parent' table and add a reference to their custodian in the Person table. This is to facilitate going from Parent to Form when the custodian, not the Parent, is the FormContact. Here is a simplified, mock structure of a sql-db I am working with. I would have drawn the relationship tables if I had access to Visio. We have a table 'Person' and we have a table 'Form'. There is a table, 'FormContact', that relates a Person to a Form, not all Persons are related to a Form. There is a relationship table for Person to Person relationships (Employer, Parent, etc.) I've asked, "Why this couldn't be handled by a query?" Response, Inefficient. (Really!?!) So, I ask, "Why not have a reference to the Form? That would be more efficient since you wouldn't be querying the FormContacts table with the reference from child/custodian." Response, this would essentially make the Parent is a FormContact. (Fair enough.) I went ahead an wrote a query to get from non-FormContact Parent to Form, and tested on the production server. The response time was instantaneous. *SOME_VALUE* is the Parent's fk ID. SELECT FormID FROM FormContact WHERE FormContact.ContactID IN (SELECT SourceContactID FROM ContactRelationship WHERE (ContactRelationship.RelatedContactID = *SOME_VALUE*) AND (ContactRelationship.Relationship = 'Parent')); If I am right, "This is an unnecessary change." What should I do, defend my position or should I concede to the managers request? If I am wrong. What is my error? Is there a better solution than the manager's?

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  • Adding rows with linq trouble with reference table

    - by Laurence Burke
    I am adding a new address and I know the structure as AddressID = PK and all other entities are non nullable. Now on insert of a new row the addrID Pk is autogened and I am wondering if I would have to get that to create a new row in the referencing table EmployeeAddress or does that automatically get generated also. also I want to be able to repopulate the dropdownlist that lists the current employee's addresses with the newly created address. static uint _curEmpID; protected void btnAdd_Click(object sender, EventArgs e) { if (txtZip.Text != "" && txtAdd1.Text != "" && txtCity.Text != "") { TestDataClassDataContext dc = new TestDataClassDataContext(); Address addr = new Address() { AddressLine1 = txtAdd1.Text, AddressLine2 = txtAdd2.Text, City = txtCity.Text, PostalCode = txtZip.Text, StateProvinceID = Convert.ToInt32(ddlState.SelectedValue) }; dc.Addresses.InsertOnSubmit(addr); lblSuccess.Visible = true; lblErrMsg.Visible = false; dc.SubmitChanges(); // // TODO: insert new row in EmployeeAddress to reference CurEmp to newly created address // SetAddrList(); } else { lblErrMsg.Text = "Invalid Input"; lblErrMsg.Visible = true; } } protected void ddlAddList_SelectedIndexChanged(object sender, EventArgs e) { lblErrMsg.Visible = false; lblSuccess.Visible = false; TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; if (ddlAddList.SelectedValue != "-1") { var addr = (from a in dc.Addresses where a.AddressID == Convert.ToInt32(ddlAddList.SelectedValue) select a).FirstOrDefault(); txtAdd1.Text = addr.AddressLine1; txtAdd2.Text = addr.AddressLine2; txtCity.Text = addr.City; txtZip.Text = addr.PostalCode; ddlState.SelectedValue = addr.StateProvinceID.ToString(); btnSubmit.Visible = true; btnAdd.Visible = false; } else { txtAdd1.Text = ""; txtAdd2.Text = ""; txtCity.Text = ""; txtZip.Text = ""; btnAdd.Visible = true; btnSubmit.Visible = false; } } protected void SetAddrList() { TestDataClassDataContext dc = new TestDataClassDataContext(); dc.ObjectTrackingEnabled = false; var addList = from addr in dc.Addresses from eaddr in dc.EmployeeAddresses where eaddr.EmployeeID == _curEmpID && addr.AddressID == eaddr.AddressID select new { AddValue = addr.AddressID, AddText = addr.AddressID, }; ddlAddList.DataSource = addList; ddlAddList.DataValueField = "AddValue"; ddlAddList.DataTextField = "AddText"; ddlAddList.DataBind(); ddlAddList.Items.Add(new ListItem("<Add Address>", "-1")); } OK I am hoping that I did not include too much code. I would really appreciate any other comments about I could otherwise improve this code in any other ways also.

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  • Multiprogramming in Django, writing to the Database

    - by Marcus Whybrow
    Introduction I have the following code which checks to see if a similar model exists in the database, and if it does not it creates the new model: class BookProfile(): # ... def save(self, *args, **kwargs): uniqueConstraint = {'book_instance': self.book_instance, 'collection': self.collection} # Test for other objects with identical values profiles = BookProfile.objects.filter(Q(**uniqueConstraint) & ~Q(pk=self.pk)) # If none are found create the object, else fail. if len(profiles) == 0: super(BookProfile, self).save(*args, **kwargs) else: raise ValidationError('A Book Profile for that book instance in that collection already exists') I first build my constraints, then search for a model with those values which I am enforcing must be unique Q(**uniqueConstraint). In addition I ensure that if the save method is updating and not inserting, that we do not find this object when looking for other similar objects ~Q(pk=self.pk). I should mention that I ham implementing soft delete (with a modified objects manager which only shows non-deleted objects) which is why I must check for myself rather then relying on unique_together errors. Problem Right thats the introduction out of the way. My problem is that when multiple identical objects are saved in quick (or as near as simultaneous) succession, sometimes both get added even though the first being added should prevent the second. I have tested the code in the shell and it succeeds every time I run it. Thus my assumption is if say we have two objects being added Object A and Object B. Object A runs its check upon save() being called. Then the process saving Object B gets some time on the processor. Object B runs that same test, but Object A has not yet been added so Object B is added to the database. Then Object A regains control of the processor, and has allready run its test, even though identical Object B is in the database, it adds it regardless. My Thoughts The reason I fear multiprogramming could be involved is that each Object A and Object is being added through an API save view, so a request to the view is made for each save, thus not a single request with multiple sequential saves on objects. It might be the case that Apache is creating a process for each request, and thus causing the problems I think I am seeing. As you would expect, the problem only occurs sometimes, which is characteristic of multiprogramming or multiprocessing errors. If this is the case, is there a way to make the test and set parts of the save() method a critical section, so that a process switch cannot happen between the test and the set?

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  • deleted gen folder, eclipse isn't generating it now :(

    - by LuxuryMode
    I accidentally deleted my gen folder and now, predictably, my resources are all messed up. I just created a gen folder myself and tried to project clean - that didn't work. Tried right-clicking project and going to android tools fix project properties - didn't work. Tried unchecking build automatically...didn't work. cleaned, closed project, closed eclipse, restarted, etc, etc. Nothing is working and I keep seeing this error: gen already exists but is not a source folder. Convert to a source folder or rename it. EDIT - OK was able to generate R.java, but now I'm getting crazy stuff in the console: [2011-06-14 17:06:11 - fastapp] Conversion to Dalvik format failed with error 1 [2011-06-14 17:06:42 - fastapp] Dx trouble processing "java/awt/font/NumericShaper.class": Ill-advised or mistaken usage of a core class (java.* or javax.*) when not building a core library. This is often due to inadvertently including a core library file in your application's project, when using an IDE (such as Eclipse). If you are sure you're not intentionally defining a core class, then this is the most likely explanation of what's going on. However, you might actually be trying to define a class in a core namespace, the source of which you may have taken, for example, from a non-Android virtual machine project. This will most assuredly not work. At a minimum, it jeopardizes the compatibility of your app with future versions of the platform. It is also often of questionable legality. If you really intend to build a core library -- which is only appropriate as part of creating a full virtual machine distribution, as opposed to compiling an application -- then use the "--core-library" option to suppress this error message. If you go ahead and use "--core-library" but are in fact building an application, then be forewarned that your application will still fail to build or run, at some point. Please be prepared for angry customers who find, for example, that your application ceases to function once they upgrade their operating system. You will be to blame for this problem. If you are legitimately using some code that happens to be in a core package, then the easiest safe alternative you have is to repackage that code. That is, move the classes in question into your own package namespace. This means that they will never be in conflict with core system classes. JarJar is a tool that may help you in this endeavor. If you find that you cannot do this, then that is an indication that the path you are on will ultimately lead to pain, suffering, grief, and lamentation. [2011-06-14 17:06:42 - fastapp] Dx 1 error; aborting [2011-06-14 17:06:42 - fastapp] Conversion to Dalvik format failed with error 1 And eclipse can't resolve the import of my resources import com.me.fastapp.R;

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  • should I ever put a major version number into a C#/Java namespace?

    - by Andrew Patterson
    I am designing a set of 'service' layer objects (data objects and interface definitions) for a WCF web service (that will be consumed by third party clients i.e. not in-house, so outside my direct control). I know that I am not going to get the interface definition exactly right - and am wanting to prepare for the time when I know that I will have to introduce a breaking set of new data objects. However, the reality of the world I am in is that I will also need to run my first version simultaneously for quite a while. The first version of my service will have URL of http://host/app/v1service.svc and when the times comes by new version will live at http://host/app/v2service.svc However, when it comes to the data objects and interfaces, I am toying with putting the 'major' version of the interface number into the actual namespace of the classes. namespace Company.Product.V1 { [DataContract(Namespace = "company-product-v1")] public class Widget { [DataMember] string widgetName; } public interface IFunction { Widget GetWidgetData(int code); } } When the time comes for a fundamental change to the service, I will introduce some classes like namespace Company.Product.V2 { [DataContract(Namespace = "company-product-v2")] public class Widget { [DataMember] int widgetCode; [DataMember] int widgetExpiry; } public interface IFunction { Widget GetWidgetData(int code); } } The advantages as I see it are that I will be able to have a single set of code serving both interface versions, sharing functionality where possible. This is because I will be able to reference both interface versions as a distinct set of C# objects. Similarly, clients may use both interface versions simultaneously, perhaps using V1.Widget in some legacy code whilst new bits move on to V2.Widget. Can anyone tell why this is a stupid idea? I have a nagging feeling that this is a bit smelly.. notes: I am obviously not proposing every single new version of the service would be in a new namespace. Presumably I will do as many non-breaking interface changes as possible, but I know that I will hit a point where all the data modelling will probably need a significant rewrite. I understand assembly versioning etc but I think this question is tangential to that type of versioning. But I could be wrong.

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  • How does the rsync algorithm correctly identify repeating blocks?

    - by Kai
    I'm on a personal quest to learn how the rsync algorithm works. After some reading and thinking, I've come up with a situation where I think the algorithm fails. I'm trying to figure out how this is resolved in an actual implementation. Consider this example, where A is the receiver and B is the sender. A = abcde1234512345fghij B = abcde12345fghij As you can see, the only change is that 12345 has been removed. Now, to make this example interesting, let's choose a block size of 5 bytes (chars). Hashing the values on the sender's side using the weak checksum gives the following values list. abcde|12345|fghij abcde -> 495 12345 -> 255 fghij -> 520 values = [495, 255, 520] Next we check to see if any hash values differ in A. If there's a matching block we can skip to the end of that block for the next check. If there's a non-matching block then we've found a difference. I'll step through this process. Hash the first block. Does this hash exist in the values list? abcde -> 495 (yes, so skip) Hash the second block. Does this hash exist in the values list? 12345 -> 255 (yes, so skip) Hash the third block. Does this hash exist in the values list? 12345 -> 255 (yes, so skip) Hash the fourth block. Does this hash exist in the values list? fghij -> 520 (yes, so skip) No more data, we're done. Since every hash was found in the values list, we conclude that A and B are the same. Which, in my humble opinion, isn't true. It seems to me this will happen whenever there is more than one block that share the same hash. What am I missing?

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  • Ruby on Rails check box not updating on form submission

    - by user284194
    I have an entries controller that allows users to add contact information the website. The user-submitted information isn't visible to users until the administrator checks a check box and submits the form. So basically my problem is that if I check the check box as an administrator while initially creating an entry (entries#new) the entry will be publicly visible as expected, but if a non-admin user creates an entry (the normal user view doesn't include the 'live' check box, only the admin one does) then that entry is stuck in limbo because the entries#edit view for some reason doesn't update the boolean check box value when logged in as an admin. entries#new view: <% form_for(@entry) do |f| %> <%= f.error_messages %> Name<br /> <%= f.text_field :name %> Mailing Address<br /> <%= f.text_field :address %> #... <%- if current_user -%> <%= f.label :live %><br /> <%= f.check_box :live %> <%- end -%> <%= f.submit 'Create' %> <% end %> entries#edit (only accessible by admin) view: <% form_for(@entry) do |f| %> <%= f.error_messages %> <%= f.label :name %><br /> <%= f.text_field :name %> Mailing Address<br /> <%= f.text_field :address %> <%= f.label :live %><br /> <%= f.check_box :live %> <%= f.submit 'Update' %> <% end %> Any ideas as to why an administrator can't update the :live check box from the edit view? I would greatly appreciate any suggestions. I'm new to rails. I can post more code if it's needed. Thanks for reading my question.

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  • QValidator for hex input

    - by Evan Teran
    I have a Qt widget which should only accept a hex string as input. It is very simple to restrict the input characters to [0-9A-Fa-f], but I would like to have it display with a delimiter between "bytes" so for example if the delimiter is a space, and the user types 0011223344 I would like the line edit to display 00 11 22 33 44 Now if the user presses the backspace key 3 times, then I want it to display 00 11 22 3. I almost have what i want, so far there is only one subtle bug involving using the delete key to remove a delimiter. Does anyone have a better way to implement this validator? Here's my code so far: class HexStringValidator : public QValidator { public: HexStringValidator(QObject * parent) : QValidator(parent) {} public: virtual void fixup(QString &input) const { QString temp; int index = 0; // every 2 digits insert a space if they didn't explicitly type one Q_FOREACH(QChar ch, input) { if(std::isxdigit(ch.toAscii())) { if(index != 0 && (index & 1) == 0) { temp += ' '; } temp += ch.toUpper(); ++index; } } input = temp; } virtual State validate(QString &input, int &pos) const { if(!input.isEmpty()) { // TODO: can we detect if the char which was JUST deleted // (if any was deleted) was a space? and special case this? // as to not have the bug in this case? const int char_pos = pos - input.left(pos).count(' '); int chars = 0; fixup(input); pos = 0; while(chars != char_pos) { if(input[pos] != ' ') { ++chars; } ++pos; } // favor the right side of a space if(input[pos] == ' ') { ++pos; } } return QValidator::Acceptable; } }; For now this code is functional enough, but I'd love to have it work 100% as expected. Obviously the ideal would be the just separate the display of the hex string from the actual characters stored in the QLineEdit's internal buffer but I have no idea where to start with that and I imagine is a non-trivial undertaking. In essence, I would like to have a Validator which conforms to this regex: "[0-9A-Fa-f]( [0-9A-Fa-f])*" but I don't want the user to ever have to type a space as delimiter. Likewise, when editing what they types, the spaces should be managed implicitly.

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  • Mocking a concrete class : templates and avoiding conditional compilation

    - by AshirusNW
    I'm trying to testing a concrete object with this sort of structure. class Database { public: Database(Server server) : server_(server) {} int Query(const char* expression) { server_.Connect(); return server_.ExecuteQuery(); } private: Server server_; }; i.e. it has no virtual functions, let alone a well-defined interface. I want to a fake database which calls mock services for testing. Even worse, I want the same code to be either built against the real version or the fake so that the same testing code can both: Test the real Database implementation - for integration tests Test the fake implementation, which calls mock services To solve this, I'm using a templated fake, like this: #ifndef INTEGRATION_TESTS class FakeDatabase { public: FakeDatabase() : realDb_(mockServer_) {} int Query(const char* expression) { MOCK_EXPECT_CALL(mockServer_, Query, 3); return realDb_.Query(); } private: // in non-INTEGRATION_TESTS builds, Server is a mock Server with // extra testing methods that allows mocking Server mockServer_; Database realDb_; }; #endif template <class T> class TestDatabaseContainer { public: int Query(const char* expression) { int result = database_.Query(expression); std::cout << "LOG: " << result << endl; return result; } private: T database_; }; Edit: Note the fake Database must call the real Database (but with a mock Server). Now to switch between them I'm planning the following test framework: class DatabaseTests { public: #ifdef INTEGRATION_TESTS typedef TestDatabaseContainer<Database> TestDatabase ; #else typedef TestDatabaseContainer<FakeDatabase> TestDatabase ; #endif TestDatabase& GetDb() { return _testDatabase; } private: TestDatabase _testDatabase; }; class QueryTestCase : public DatabaseTests { public: void TestStep1() { ASSERT(GetDb().Query(static_cast<const char *>("")) == 3); return; } }; I'm not a big fan of that compile-time switching between the real and the fake. So, my question is: Whether there's a better way of switching between Database and FakeDatabase? For instance, is it possible to do it at runtime in a clean fashion? I like to avoid #ifdefs. Also, if anyone has a better way of making a fake class that mimics a concrete class, I'd appreciate it. I don't want to have templated code all over the actual test code (QueryTestCase class). Feel free to critique the code style itself, too. You can see a compiled version of this code on codepad.

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  • How to know when a user has really released a key in Java?

    - by Luis Soeiro
    (Edited for clarity) I want to detect when a user presses and releases a key in Java Swing, ignoring the keyboard auto repeat feature. I also would like a pure Java approach the works on Linux, Mac OS and Windows. Requirements: When the user presses some key I want to know what key is that; When the user releases some key, I want to know what key is that; I want to ignore the system auto repeat options: I want to receive just one keypress event for each key press and just one key release event for each key release; If possible, I would use items 1 to 3 to know if the user is holding more than one key at a time (i.e, she hits 'a' and without releasing it, she hits "Enter"). The problem I'm facing in Java is that under Linux, when the user holds some key, there are many keyPress and keyRelease events being fired (because of the keyboard repeat feature). I've tried some approaches with no success: Get the last time a key event occurred - in Linux, they seem to be zero for key repeat, however, in Mac OS they are not; Consider an event only if the current keyCode is different from the last one - this way the user can't hit twice the same key in a row; Here is the basic (non working) part of code: import java.awt.event.KeyListener; public class Example implements KeyListener { public void keyTyped(KeyEvent e) { } public void keyPressed(KeyEvent e) { System.out.println("KeyPressed: "+e.getKeyCode()+", ts="+e.getWhen()); } public void keyReleased(KeyEvent e) { System.out.println("KeyReleased: "+e.getKeyCode()+", ts="+e.getWhen()); } } When a user holds a key (i.e, 'p') the system shows: KeyPressed: 80, ts=1253637271673 KeyReleased: 80, ts=1253637271923 KeyPressed: 80, ts=1253637271923 KeyReleased: 80, ts=1253637271956 KeyPressed: 80, ts=1253637271956 KeyReleased: 80, ts=1253637271990 KeyPressed: 80, ts=1253637271990 KeyReleased: 80, ts=1253637272023 KeyPressed: 80, ts=1253637272023 ... At least under Linux, the JVM keeps resending all the key events when a key is being hold. To make things more difficult, on my system (Kubuntu 9.04 Core 2 Duo) the timestamps keep changing. The JVM sends a key new release and new key press with the same timestamp. This makes it hard to know when a key is really released. Any ideas? Thanks

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  • Consolidating coding styles: Funcs, private method, single method classes

    - by jdoig
    Hi all, We currently have 3 devs with, some, conflicting styles and I'm looking for a way to bring peace to the kingdom... The Coders: Foo 1: Likes to use Func's & Action's inside public methods. He uses actions to alias off lengthy method calls and Func's to perform simple tasks that can be expressed in 1 or 2 lines and will be used frequently through out the code Pros: The main body of his code is succinct and very readable, often with only one or 2 public methods per class and rarely any private methods. Cons: The start of methods contain blocks of lambda rich code that other developers don't enjoy reading; and, on occasion, can contain higher order functions that other dev's REALLY don't like reading. Foo 2: Likes to create a private method for (almost) everything the public method will have to do . Pros: Public methods remain small and readable (to all developers). Cons: Private methods are numerous. With private methods that call into other private methods, that call into... etc, etc. Making code hard to navigate. Foo 3: Likes to create a public class with a, single, public method for every, non-trivial, task that needs performing, then dependency inject them into other objects. Pros: Easily testable, easy to understand (one object, one responsibility). Cons: project gets littered by classes, opening multiple class files to understand what code does makes navigation awkward. It would be great to take the best of all these techniques... Foo-1 Has really nice, readable (almost dsl-like) code... for the most part, except for all the Action and Func lambda shenanigans bulked together at the start of a method. Foo-3 Has highly testable and extensible code that just feels a bit "belt-&-braces" for some solutions and has some code-navigation niggles (constantly hitting F12 in VS and opening 5 other .cs files to find out what a single method does). And Foo-2... Well I'm not sure I like anything about the one-huge .cs file with 2 public methods and 12 private ones, except for the fact it's easier for juniors to dig into. I admit I grossly over-simplified the explanations of those coding styles; but if any one knows of any patterns, practices or diplomatic-manoeuvres that can help unite our three developers (without just telling any of them to just "stop it!") that would be great. From a feasibility standpoint : Foo-1's style meets with the most resistance due to some developers finding lambda and/or Func's hard to read. Foo-2's style meets with a less resistance as it's just so easy to fall into. Foo-3's style requires the most forward thinking and is difficult to enforce when time is short. Any ideas on some coding styles or conventions that can make this work?

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  • Spring.NET & Immediacy CMS (or how to inject to server side controls without using PageHandlerFactor

    - by Simon Rice
    Is there any way to inject dependencies into an Immediacy CMS control using Spring.NET, ideally without having to use to ContextRegistry when initialising the control? Update, with my own answer The issue here is that Immediacy already has a handler defined in web.config that deals with all aspx pages, & so it's not possible add an entry for Spring.NET's PageHandlerFactory in web.config as per a normal webforms app. That rules out making the control implement ISupportsWebDependencyInjection. Furthermore, most of Immediacy's generated pages are aspx pages that don't physically exist on the drive. I have changed the title of the question to reflect this. What I have done to get Dependency Injection working is: Add the usual entries to web.config for Spring.NET as outlined in the documentation, except for the adding the entry to the <httpHandlers> section. In this case I've got my object definitions in Spring.config. Create the following abstract base class that will deal with all of the Dependency Injection work: DIControl.cs public abstract class DIControl : ImmediacyControl { protected virtual string DIName { get { return this.GetType().Name; } } protected override void OnInit(EventArgs e) { if (ContextRegistry.GetContext().GetObject(DIName, this.GetType()) != null) ContextRegistry.GetContext().ConfigureObject(this, DIName); base.OnInit(e); } } For non-immediacy controls, you can make this server side control inherit from Control or whatever subclass of that you like. For any control with which you wish to use with Spring.NET's Inversion of Control container, define it to inherit from DIControl & add the relelvant entry to Spring.config, for example: SampleControl.cs public class SampleControl : DIControl, INamingContainer { public string Text { get; set; } protected string InjectedText { get; set; } public SampleControl() : base() { Text = "Hello world"; } protected override void RenderContents(HtmlTextWriter output) { output.Write(string.Format("{0} {1}", Text, InjectedText)); } } Spring.config <objects xmlns="http://www.springframework.net"> <object id="SampleControl" type="MyProject.SampleControl, MyAssembly"> <property name="InjectedText" value="from Spring.NET" /> </object> </objects> You can optionally override DIName if you wish to name your entry in Spring.config differently from the name of your class. Provided everything's done correctly, you will have the control writing out "Hello world from Spring.NET!" when used in a page. This solution uses Spring.NET's ContextRegistry from within the control, but I would be surprised if there's no way around that for Immediacy at least since the page objects themselves aren't accessible. However, can this be improved at all from a Spring.NET perspective? Is there maybe an Immediacy plugin that already does this that I'm completely unaware of? Or is there an approach that does this in a more elegant way? I'm open to suggestions.

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  • Make a compiled binary run at native speed flawlessly without recompiling from source on a another system?

    - by unknownthreat
    I know that many people, at a first glance of the question, may immediately yell out "Java", but no, I know Java's qualities. Allow me to elaborate my question first. Normally, when we want our program to run at a native speed on a system, whether it be Windows, Mac OS X, or Linux, we need to compile from source codes. If you want to run a program of another system in your system, you need to use a virtual machine or an emulator. While these tools allow you to use the program you need on the non-native OS, they sometimes have problems of performance and glitches. We also have a newer compiler called "JIT Compiler", where the compiler will parse the bytecode program to native machine language before execution. The performance may increase to a very good extent with JIT Compiler, but the performance is still not the same as running it on a native system. Another program on Linux, WINE, is also a good tool for running Windows program on Linux system. I have tried running Team Fortress 2 on it, and tried experiment with some settings. I got ~40 fps on Windows at its mid-high setting on 1280 x 1024. On Linux, I need to turn everything low at 1280 x 1024 to get ~40 fps. There are 2 notable things though: Polygon model settings do not seem to affect framerate whether I set it low or high. When there are post-processing effects or some special effects that require manipulation of drawn pixels of the current frame, the framerate will drop to 10-20 fps. From this point, I can see that normal polygon rendering is just fine, but when it comes to newer rendering methods that requires graphic card to the job, it slows down to a crawl. Anyway, this question is rather theoretical. Is there anything we can do at all? I see that WINE can run STEAM and Team Fortress 2. Although there are flaws, they can run at lower setting. Or perhaps, I should also ask, "is it possible to translate one whole program on a system to another system without recompiling from source and get native speed?" I see that we also have AOT Compiler, is it possible to use it for something like this? Or there are so many constraints (such as DirectX call or differences in software architecture) that make it impossible to have a flawless and not native to the system program that runs at native speed?

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  • Internal class and access to external members.

    - by Knowing me knowing you
    I always thought that internal class has access to all data in its external class but having code: template<class T> class Vector { template<class T> friend std::ostream& operator<<(std::ostream& out, const Vector<T>& obj); private: T** myData_; std::size_t myIndex_; std::size_t mySize_; public: Vector():myData_(nullptr), myIndex_(0), mySize_(0) { } Vector(const Vector<T>& pattern); void insert(const T&); Vector<T> makeUnion(const Vector<T>&)const; Vector<T> makeIntersection(const Vector<T>&)const; class Iterator : public std::iterator<std::bidirectional_iterator_tag,T> { private: T** itData_; public: Iterator()//<<<<<<<<<<<<<------------COMMENT { /*HERE I'M TRYING TO USE ANY MEMBER FROM Vector<T> AND I'M GETTING ERR SAYING: ILLEGAL CALL OF NON-STATIC MEMBER FUNCTION*/} Iterator(T** ty) { itData_ = ty; } Iterator operator++() { return ++itData_; } T operator*() { return *itData_[0]; } bool operator==(const Iterator& obj) { return *itData_ == *obj.itData_; } bool operator!=(const Iterator& obj) { return *itData_ != *obj.itData_; } bool operator<(const Iterator& obj) { return *itData_ < *obj.itData_; } }; typedef Iterator iterator; iterator begin()const { assert(mySize_ > 0); return myData_; } iterator end()const { return myData_ + myIndex_; } }; See line marked as COMMENT. So can I or I can't use members from external class while in internal class? Don't bother about naming, it's not a Vector it's a Set. Thank you.

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  • What is the difference between NULL in C++ and null in Java?

    - by Stephano
    I've been trying to figure out why C++ is making me crazy typing NULL. Suddenly it hits me the other day; I've been typing null (lower case) in Java for years. Now suddenly I'm programming in C++ and that little chunk of muscle memory is making me crazy. Wikiperipatetic defines C++ NULL as part of the stddef: A macro that expands to a null pointer constant. It may be defined as ((void*)0), 0 or 0L depending on the compiler and the language. Sun's docs tells me this about Java's "null literal": The null type has one value, the null reference, represented by the literal null, which is formed from ASCII characters. A null literal is always of the null type. So this is all very nice. I know what a null pointer reference is, and thank you for the compiler notes. Now I'm a little fuzzy on the idea of a literal in Java so I read on... A literal is the source code representation of a fixed value; literals are represented directly in your code without requiring computation. There's also a special null literal that can be used as a value for any reference type. null may be assigned to any variable, except variables of primitive types. There's little you can do with a null value beyond testing for its presence. Therefore, null is often used in programs as a marker to indicate that some object is unavailable. Ok, so I think I get it now. In C++ NULL is a macro that, when compiled, defines the null pointer constant. In Java, null is a fixed value that any non-primitive can be assigned too; great for testing in a handy if statement. Java does not have pointers, so I can see why they kept null a simple value rather than anything fancy. But why did java decide to change the all caps NULL to null? Furthermore, am I missing anything here?

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  • How can I structure and recode messy categorical data in R?

    - by briandk
    I'm struggling with how to best structure categorical data that's messy, and comes from a dataset I'll need to clean. The Coding Scheme I'm analyzing data from a university science course exam. We're looking at patterns in student responses, and we developed a coding scheme to represent the kinds of things students are doing in their answers. A subset of the coding scheme is shown below. Note that within each major code (1, 2, 3) are nested non-unique sub-codes (a, b, ...). What the Raw Data Looks Like I've created an anonymized, raw subset of my actual data which you can view here. Part of my problem is that those who coded the data noticed that some students displayed multiple patterns. The coders' solution was to create enough columns (reason1, reason2, ...) to hold students with multiple patterns. That becomes important because the order (reason1, reason2) is arbitrary--two students (like student 41 and student 42 in my dataset) who correctly applied "dependency" should both register in an analysis, regardless of whether 3a appears in the reason column or the reason2 column. How Can I Best Structure Student Data? Part of my problem is that in the raw data, not all students display the same patterns, or the same number of them, in the same order. Some students may do just one thing, others may do several. So, an abstracted representation of example students might look like this: Note in the example above that student002 and student003 both are coded as "1b", although I've deliberately shown the order as different to reflect the reality of my data. My (Practical) Questions Should I concatenate reason1, reason2, ... into one column? How can I (re)code the reasons in R to reflect the multiplicity for some students? Thanks I realize this question is as much about good data conceptualization as it is about specific features of R, but I thought it would be appropriate to ask it here. If you feel it's inappropriate for me to ask the question, please let me know in the comments, and stackoverflow will automatically flood my inbox with sadface emoticons. If I haven't been specific enough, please let me know and I'll do my best to be clearer.

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  • E4X in ActionScript help needed

    - by voipsecuritydigest.com
    Here is the XML How using E4X read values of nodes <status>??</status> and of node <invisible value="false"/> ? <?xml version="1.0" encoding="utf-8"?> <s:Application xmlns:fx="http://ns.adobe.com/mxml/2009" xmlns:s="library://ns.adobe.com/flex/spark" creationComplete="init()"> <fx:Declarations> <!-- Place non-visual elements (e.g., services, value objects) here --> </fx:Declarations> <fx:Script> <![CDATA[ var xml:XML = <iq type="result" id="ss-1"> <query status-min-ver="1" status-max="512" status-list-contents-max="5" status-list-max="3" xmlns="google:shared-status"> <status> ?? </status> <show> default </show> <status-list show="default"> <status> ?? </status> <status> ? </status> <status> ?? </status> </status-list> <status-list show="dnd"> <status> ?? </status> <status> dnd, i have bad mood </status> <status> showering </status> <status> ??_???¦ </status> <status> ? </status> </status-list> <invisible value="false"/> </query> </iq> public function init() { trace(xml.query.invisible.@value); } ]]> </fx:Script> </s:Application>

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  • How to troubleshoot a Highcharts script that's not rendering data when date is added and hanging the JS engine with large datasets?

    - by ylluminate
    I have a Highchart JS graph that I'm building in Rails (although I don't think Ruby has real bearing on this problem unless it's the Date output format) to which I'm adding the timestamp of each datapoint. Presently the array of floats is rendering fine without timestamps, however when I add the timestamp to the series it fails to rend. What's worse is that when the series has hundreds of entries all sorts of problems arise, not the least of which is the browser entirely hanging and requiring a force quit / kill. I'm using the following to build the array of arrays data series: series1 = readings.map{|row| [(row.date.to_i * 1000), (row.data1.to_f if BigDecimal(row.data1) != BigDecimal("-1000.0"))] } This yields a result like this: series: [{"name":"Data 1","data":[[1326262980000,1.79e-09],[1326262920000,1.29e-09],[1326262860000,1.22e-09],[1326262800000,1.42e-09],[1326262740000,1.29e-09],[1326262680000,1.34e-09],[1326262620000,1.31e-09],[1326262560000,1.51e-09],[1326262500000,1.24e-09],[1326262440000,1.7e-09],[1326262380000,1.24e-09],[1326262320000,1.29e-09],[1326262260000,1.53e-09],[1326262200000,1.23e-09],[1326262140000,1.21e-09]],"color":"blue"}] Yet nothing appears on the graph as noted. Notwithstanding, when I compare the data series in one of their very similar examples here: http://www.highcharts.com/demo/spline-irregular-time It appears that really the data series are formatted identically (except in mine I use the timestamp vs date method). This leads me to think I've got a problem with the timestamp output, but I'm just not able to figure out where / how as I'm converting the date output to an integer multipled by 1000 to convert it to milliseconds as per explained in a similar Railscasts tutorial. I would very much appreciate it if someone could point me in the right direction here as to what I may be doing wrong. What could cause no data to appear on the graph in smaller sized sets (<100 points) and when into the hundreds causes an apparent hang in the javascript engine in this case? Perhaps ultimately the key lies here as this is the entire js that's being generated and not rendering: jQuery(function() { // 1. Define JSON options var options = { chart: {"defaultSeriesType":"spline","renderTo":"chart_name"}, title: {"text":"Title"}, legend: {"layout":"vertical","style":{}}, xAxis: {"title":{"text":"UTC Time"},"type":"datetime"}, yAxis: [{"title":{"text":"Left Title","margin":10}},{"title":{"text":"Right Groups Title"},"opposite":true}], tooltip: {"enabled":true}, credits: {"enabled":false}, plotOptions: {"areaspline":{}}, series: [{"name":"Data 1","data":[[1326262980000,1.79e-08],[1326262920000,1.69e-08],[1326262860000,1.62e-08],[1326262800000,1.42e-08],[1326262740000,1.29e-08],[1326262680000,1.34e-08],[1326262620000,1.31e-08],[1326262560000,1.51e-08],[1326262500000,1.64e-08],[1326262440000,1.7e-08],[1326262380000,1.64e-08],[1326262320000,1.69e-08],[1326262260000,1.53e-08],[1326262200000,1.23e-08],[1326262140000,1.21e-08]],"color":"blue"},{"name":"Data 2","data":[[1326262980000,9.79e-09],[1326262920000,9.78e-09],[1326262860000,9.8e-09],[1326262800000,9.82e-09],[1326262740000,9.88e-09],[1326262680000,9.89e-09],[1326262620000,1.3e-06],[1326262560000,1.32e-06],[1326262500000,1.33e-06],[1326262440000,1.33e-06],[1326262380000,1.34e-06],[1326262320000,1.33e-06],[1326262260000,1.32e-06],[1326262200000,1.32e-06],[1326262140000,1.32e-06]],"color":"red"}], subtitle: {} }; // 2. Add callbacks (non-JSON compliant) // 3. Build the chart var chart = new Highcharts.StockChart(options); });

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