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  • JavaFX FXML communication between Application and Controller classes

    - by likethesky
    I am trying to get and destroy an external process I've created via ProcessBuilder in my FXML application close, but it's not working. This is based on the helpful advice Sergey Grinev gave me here. I have tried running with/without the "// myController.setApp(this);" and with "// super.stop();" at top of subclass and at bottom (see commented out/in for that line in MyApp), but no combination works. This probably isn't related to FXML or JavaFX, though I imagine this is a common pattern for developing apps on JavaFX. I suppose I'm asking for a Java best practice for closing dependent processes in a UI-based app like this one (in this case: FXML / JavaFX based), where there is a controller class and an application class. Can you explain what I'm doing wrong? Or better: advise what I should be doing instead? Thanks. In my Application I do this: public class MyApp extends Application { @Override public void start(Stage primaryStage) throws Exception { FXMLLoader fxmlLoader = new FXMLLoader(); Scene scene = (Scene)FXMLLoader.load(getClass().getResource("MyApp.fxml")); MyAppController myController = (MyAppController)fxmlLoader.getController(); primaryStage.setScene(scene); primaryStage.show(); // myController.setApp(this); } @Override public void stop() throws Exception { // super.stop(); // this is called on fx app close, you may call it in an action handler too if (MyAppController.getScriptProcess() != null) { MyAppController.getScriptProcess().destroy(); } super.stop(); } public static void main(String[] args) { launch(args); } } In my Controller I do this: public class MyAppController implements Initializable { private Application app; private static Process scriptProcess; public void setApp(Application a) { app = a; } public static Process getScriptProcess() { return scriptProcess; } } The result when I run with the "commented-out setApp()" not commented out (that is, left in the start method), is the following, immediately upon launch (the main Scene flashes, then disappears, then this dialog appears: "JavaFX Launcher Error: Exception while running Application" And it gives an, "Exception in Application start method" in the console as well. The result when I leave out the "commented-out code" in my MyApp above (that is, remove the "setApp()" from the start method), is that my app does indeed close, but gives this error when it closes: Exception in thread "JavaFX Application Thread" java.lang.RuntimeException: java.lang.reflect.InvocationTargetException at javafx.fxml.FXMLLoader$ControllerMethodEventHandler.handle(FXMLLoader.java:1440) at com.sun.javafx.event.CompositeEventHandler.dispatchBubblingEvent(CompositeEventHandler.java:69) at com.sun.javafx.event.EventHandlerManager.dispatchBubblingEvent(EventHandlerManager.java:217) at com.sun.javafx.event.EventHandlerManager.dispatchBubblingEvent(EventHandlerManager.java:170) at com.sun.javafx.event.CompositeEventDispatcher.dispatchBubblingEvent(CompositeEventDispatcher.java:38) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:37) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:35) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:35) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.EventUtil.fireEventImpl(EventUtil.java:53) at com.sun.javafx.event.EventUtil.fireEvent(EventUtil.java:28) at javafx.event.Event.fireEvent(Event.java:171) at javafx.scene.Node.fireEvent(Node.java:6863) at javafx.scene.control.Button.fire(Button.java:179) at com.sun.javafx.scene.control.behavior.ButtonBehavior.mouseReleased(ButtonBehavior.java:193) at com.sun.javafx.scene.control.skin.SkinBase$4.handle(SkinBase.java:336) at com.sun.javafx.scene.control.skin.SkinBase$4.handle(SkinBase.java:329) at com.sun.javafx.event.CompositeEventHandler.dispatchBubblingEvent(CompositeEventHandler.java:64) at com.sun.javafx.event.EventHandlerManager.dispatchBubblingEvent(EventHandlerManager.java:217) at com.sun.javafx.event.EventHandlerManager.dispatchBubblingEvent(EventHandlerManager.java:170) at com.sun.javafx.event.CompositeEventDispatcher.dispatchBubblingEvent(CompositeEventDispatcher.java:38) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:37) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:35) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:35) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.BasicEventDispatcher.dispatchEvent(BasicEventDispatcher.java:35) at com.sun.javafx.event.EventDispatchChainImpl.dispatchEvent(EventDispatchChainImpl.java:92) at com.sun.javafx.event.EventUtil.fireEventImpl(EventUtil.java:53) at com.sun.javafx.event.EventUtil.fireEvent(EventUtil.java:33) at javafx.event.Event.fireEvent(Event.java:171) at javafx.scene.Scene$MouseHandler.process(Scene.java:3324) at javafx.scene.Scene$MouseHandler.process(Scene.java:3164) at javafx.scene.Scene$MouseHandler.access$1900(Scene.java:3119) at javafx.scene.Scene.impl_processMouseEvent(Scene.java:1559) at javafx.scene.Scene$ScenePeerListener.mouseEvent(Scene.java:2261) at com.sun.javafx.tk.quantum.GlassViewEventHandler.handleMouseEvent(GlassViewEventHandler.java:228) at com.sun.glass.ui.View.handleMouseEvent(View.java:528) at com.sun.glass.ui.View.notifyMouse(View.java:922) at com.sun.glass.ui.gtk.GtkApplication._runLoop(Native Method) at com.sun.glass.ui.gtk.GtkApplication$3$1.run(GtkApplication.java:82) at java.lang.Thread.run(Thread.java:722) Caused by: java.lang.reflect.InvocationTargetException at sun.reflect.NativeMethodAccessorImpl.invoke0(Native Method) at sun.reflect.NativeMethodAccessorImpl.invoke(NativeMethodAccessorImpl.java:57) at sun.reflect.DelegatingMethodAccessorImpl.invoke(DelegatingMethodAccessorImpl.java:43) at java.lang.reflect.Method.invoke(Method.java:601) at javafx.fxml.FXMLLoader$ControllerMethodEventHandler.handle(FXMLLoader.java:1435) ... 44 more Caused by: java.lang.NullPointerException at mypackage.MyController.handleCancel(MyController.java:300) ... 49 more Clean up...

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  • Help with MVC controller: passing a string from view to controller

    - by 109221793
    Hi guys, I'm having trouble with one particular issue, I was hoping someone could help me out. I've completed the MVC Music Store tutorial, and now I'm trying to add some administrator functionality - practice as I will have to do this in an MVC application in my job. The application is using the aspnet membership api, and what I have done so far is created a view to list the users. What I want to be able to do, is click on the users name in order to change their password. To try and carry the username to the changeUserPassword controller (custom made). I registered a new route in the global.asax.cs file in order to display the username in the URL, which is working so far. UserList View <%: Html.RouteLink(user.UserName, "AdminPassword", new { controller="StoreManager", action="changeUserPassword", username = user.UserName }) %> Global.asax.cs routes.MapRoute( "AdminPassword", //Route name "{controller}/{action}/{username}", //URL with parameters new { controller = "StoreManager", action = "changeUserPassword", username = UrlParameter.Optional} ); So now the URL looks like this when I reach the changeUserPassword view: http://localhost:51236/StoreManager/changeUserPassword/Administrator Here is the GET changeUserPassword action: public ActionResult changeUserPassword(string username) { ViewData["username"] = username; return View(); } I wanted to store the username in ViewData as I would like to use it in the GET changeUserPassword for display purposes, and also as a hidden value in the form. This is in order to pass it through to enable me to reset the password. Having debugged through the code, it seems that 'username' is null. How can I get this to work so that the username carries over from the Html.RouteLink, to the changeUserPassword action? Any help would be appreciated :) Here is my complete code: UserList.aspx <%@ Page Title="" Language="C#" MasterPageFile="~/Views/Shared/Site.Master" Inherits="System.Web.Mvc.ViewPage<System.Web.Security.MembershipUserCollection>" %> <asp:Content ID="Content1" ContentPlaceHolderID="TitleContent" runat="server"> UserList </asp:Content> <asp:Content ID="Content2" ContentPlaceHolderID="MainContent" runat="server"> <h2>UserList</h2> <table> <tr> <th>User Name</th> <th>Last Activity date</th> <th>Locked Out</th> </tr> <%foreach (MembershipUser user in Model){ %> <tr> <td><%: Html.RouteLink(user.UserName, "AdminPassword", new { controller="StoreManager", action="changeUserPassword", username = user.UserName }) %></td> <td><%: user.LastActivityDate %></td> <td><%: user.IsLockedOut %></td> </tr> <% }%> </table> </asp:Content> changeUserPassword.aspx <%@ Page Title="" Language="C#" MasterPageFile="~/Views/Shared/Site.Master" Inherits="System.Web.Mvc.ViewPage<musicStoreMVC.ViewModels.ResetPasswordAdmin>" %> <asp:Content ID="Content1" ContentPlaceHolderID="TitleContent" runat="server"> changeUserPassword </asp:Content> <asp:Content ID="Content2" ContentPlaceHolderID="MainContent" runat="server"> <h2>Change Password: <%: ViewData["username"] %></h2> <% using (Html.BeginForm()) {%> <%: Html.ValidationSummary(true) %> <fieldset> <legend>Fields</legend> <div class="editor-label"> <%: Html.Hidden("username",ViewData["username"]) %> <%: Html.LabelFor(model => model.password) %> </div> <div class="editor-field"> <%: Html.TextBoxFor(model => model.password) %> <%: Html.ValidationMessageFor(model => model.password) %> </div> <div class="editor-label"> <%: Html.LabelFor(model => model.confirmPassword) %> </div> <div class="editor-field"> <%: Html.TextBoxFor(model => model.confirmPassword) %> <%: Html.ValidationMessageFor(model => model.confirmPassword) %> </div> <p> <input type="submit" value="Create" /> </p> </fieldset> <% } %> <div> <%: Html.ActionLink("Back to List", "Index") %> </div> </asp:Content> My actions public ActionResult UserList() { var users = Membership.GetAllUsers(); return View(users); } public ActionResult changeUserPassword(string username) { ViewData["username"] = username; return View(); }

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  • Confusion testing fftw3 - poisson equation 2d test

    - by user3699736
    I am having trouble explaining/understanding the following phenomenon: To test fftw3 i am using the 2d poisson test case: laplacian(f(x,y)) = - g(x,y) with periodic boundary conditions. After applying the fourier transform to the equation we obtain : F(kx,ky) = G(kx,ky) /(kx² + ky²) (1) if i take g(x,y) = sin (x) + sin(y) , (x,y) \in [0,2 \pi] i have immediately f(x,y) = g(x,y) which is what i am trying to obtain with the fft : i compute G from g with a forward Fourier transform From this i can compute the Fourier transform of f with (1). Finally, i compute f with the backward Fourier transform (without forgetting to normalize by 1/(nx*ny)). In practice, the results are pretty bad? (For instance, the amplitude for N = 256 is twice the amplitude obtained with N = 512) Even worse, if i try g(x,y) = sin(x)*sin(y) , the curve has not even the same form of the solution. (note that i must change the equation; i divide by two the laplacian in this case : (1) becomes F(kx,ky) = 2*G(kx,ky)/(kx²+ky²) Here is the code: /* * fftw test -- double precision */ #include <iostream> #include <stdio.h> #include <stdlib.h> #include <math.h> #include <fftw3.h> using namespace std; int main() { int N = 128; int i, j ; double pi = 3.14159265359; double *X, *Y ; X = (double*) malloc(N*sizeof(double)); Y = (double*) malloc(N*sizeof(double)); fftw_complex *out1, *in2, *out2, *in1; fftw_plan p1, p2; double L = 2.*pi; double dx = L/((N - 1)*1.0); in1 = (fftw_complex*) fftw_malloc(sizeof(fftw_complex)*(N*N) ); out2 = (fftw_complex*) fftw_malloc(sizeof(fftw_complex)*(N*N) ); out1 = (fftw_complex*) fftw_malloc(sizeof(fftw_complex)*(N*N) ); in2 = (fftw_complex*) fftw_malloc(sizeof(fftw_complex)*(N*N) ); p1 = fftw_plan_dft_2d(N, N, in1, out1, FFTW_FORWARD,FFTW_MEASURE ); p2 = fftw_plan_dft_2d(N, N, in2, out2, FFTW_BACKWARD,FFTW_MEASURE); for(i = 0; i < N; i++){ X[i] = -pi + (i*1.0)*2.*pi/((N - 1)*1.0) ; for(j = 0; j < N; j++){ Y[j] = -pi + (j*1.0)*2.*pi/((N - 1)*1.0) ; in1[i*N + j][0] = sin(X[i]) + sin(Y[j]) ; // row major ordering //in1[i*N + j][0] = sin(X[i]) * sin(Y[j]) ; // 2nd test case in1[i*N + j][1] = 0 ; } } fftw_execute(p1); // FFT forward for ( i = 0; i < N; i++){ // f = g / ( kx² + ky² ) for( j = 0; j < N; j++){ in2[i*N + j][0] = out1[i*N + j][0]/ (i*i+j*j+1e-16); in2[i*N + j][1] = out1[i*N + j][1]/ (i*i+j*j+1e-16); //in2[i*N + j][0] = 2*out1[i*N + j][0]/ (i*i+j*j+1e-16); // 2nd test case //in2[i*N + j][1] = 2*out1[i*N + j][1]/ (i*i+j*j+1e-16); } } fftw_execute(p2); //FFT backward // checking the results computed double erl1 = 0.; for ( i = 0; i < N; i++) { for( j = 0; j < N; j++){ erl1 += fabs( in1[i*N + j][0] - out2[i*N + j][0]/N/N )*dx*dx; cout<< i <<" "<< j<<" "<< sin(X[i])+sin(Y[j])<<" "<< out2[i*N+j][0]/N/N <<" "<< endl; // > output } } cout<< erl1 << endl ; // L1 error fftw_destroy_plan(p1); fftw_destroy_plan(p2); fftw_free(out1); fftw_free(out2); fftw_free(in1); fftw_free(in2); return 0; } I can't find any (more) mistakes in my code (i installed the fftw3 library last week) and i don't see a problem with the maths either but i don't think it's the fft's fault. Hence my predicament. I am all out of ideas and all out of google as well. Any help solving this puzzle would be greatly appreciated. note : compiling : g++ test.cpp -lfftw3 -lm executing : ./a.out output and i use gnuplot in order to plot the curves : (in gnuplot ) splot "output" u 1:2:4 ( for the computed solution )

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  • Making swap faster, easier to use and exception-safe

    - by FredOverflow
    I could not sleep last night and started thinking about std::swap. Here is the familiar C++98 version: template <typename T> void swap(T& a, T& b) { T c(a); a = b; b = c; } If a user-defined class Foo uses external ressources, this is inefficient. The common idiom is to provide a method void Foo::swap(Foo& other) and a specialization of std::swap<Foo>. Note that this does not work with class templates since you cannot partially specialize a function template, and overloading names in the std namespace is illegal. The solution is to write a template function in one's own namespace and rely on argument dependent lookup to find it. This depends critically on the client to follow the "using std::swap idiom" instead of calling std::swap directly. Very brittle. In C++0x, if Foo has a user-defined move constructor and a move assignment operator, providing a custom swap method and a std::swap<Foo> specialization has little to no performance benefit, because the C++0x version of std::swap uses efficient moves instead of copies: #include <utility> template <typename T> void swap(T& a, T& b) { T c(std::move(a)); a = std::move(b); b = std::move(c); } Not having to fiddle with swap anymore already takes a lot of burden away from the programmer. Current compilers do not generate move constructors and move assignment operators automatically yet, but as far as I know, this will change. The only problem left then is exception-safety, because in general, move operations are allowed to throw, and this opens up a whole can of worms. The question "What exactly is the state of a moved-from object?" complicates things further. Then I was thinking, what exactly are the semantics of std::swap in C++0x if everything goes fine? What is the state of the objects before and after the swap? Typically, swapping via move operations does not touch external resources, only the "flat" object representations themselves. So why not simply write a swap template that does exactly that: swap the object representations? #include <cstring> template <typename T> void swap(T& a, T& b) { unsigned char c[sizeof(T)]; memcpy( c, &a, sizeof(T)); memcpy(&a, &b, sizeof(T)); memcpy(&b, c, sizeof(T)); } This is as efficient as it gets: it simply blasts through raw memory. It does not require any intervention from the user: no special swap methods or move operations have to be defined. This means that it even works in C++98 (which does not have rvalue references, mind you). But even more importantly, we can now forget about the exception-safety issues, because memcpy never throws. I can see two potential problems with this approach: First, not all objects are meant to be swapped. If a class designer hides the copy constructor or the copy assignment operator, trying to swap objects of the class should fail at compile-time. We can simply introduce some dead code that checks whether copying and assignment are legal on the type: template <typename T> void swap(T& a, T& b) { if (false) // dead code, never executed { T c(a); // copy-constructible? a = b; // assignable? } unsigned char c[sizeof(T)]; std::memcpy( c, &a, sizeof(T)); std::memcpy(&a, &b, sizeof(T)); std::memcpy(&b, c, sizeof(T)); } Any decent compiler can trivially get rid of the dead code. (There are probably better ways to check the "swap conformance", but that is not the point. What matters is that it's possible). Second, some types might perform "unusual" actions in the copy constructor and copy assignment operator. For example, they might notify observers of their change. I deem this a minor issue, because such kinds of objects probably should not have provided copy operations in the first place. Please let me know what you think of this approach to swapping. Would it work in practice? Would you use it? Can you identify library types where this would break? Do you see additional problems? Discuss!

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  • How to implement Cocoa copyWithZone on derived object in MonoMac C#?

    - by Justin Aquadro
    I'm currently porting a small Winforms-based .NET application to use a native Mac front-end with MonoMac. The application has a TreeControl with icons and text, which does not exist out of the box in Cocoa. So far, I've ported almost all of the ImageAndTextCell code in Apple's DragNDrop example: https://developer.apple.com/library/mac/#samplecode/DragNDropOutlineView/Listings/ImageAndTextCell_m.html#//apple_ref/doc/uid/DTS40008831-ImageAndTextCell_m-DontLinkElementID_6, which is assigned to an NSOutlineView as a custom cell. It seems to be working almost perfectly, except that I have not figured out how to properly port the copyWithZone method. Unfortunately, this means the internal copies that NSOutlineView is making do not have the image field, and it leads to the images briefly vanishing during expand and collapse operations. The objective-c code in question is: - (id)copyWithZone:(NSZone *)zone { ImageAndTextCell *cell = (ImageAndTextCell *)[super copyWithZone:zone]; // The image ivar will be directly copied; we need to retain or copy it. cell->image = [image retain]; return cell; } The first line is what's tripping me up, as MonoMac does not expose a copyWithZone method, and I don't know how to otherwise call it. Update Based on current answers and additional research and testing, I've come up with a variety of models for copying an object. static List<ImageAndTextCell> _refPool = new List<ImageAndTextCell>(); // Method 1 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 2 [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); return cell; } [Export("dealloc")] public void Dealloc () { _refPool.Remove(this); this.Dispose(); } // Method 3 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 4 static IntPtr selRetain = Selector.GetHandle ("retain"); static IntPtr selRetainCount = Selector.GetHandle("retainCount"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone (IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell () { Title = Title, Image = Image, }; _refPool.Add (cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } public void PeriodicCleanup () { List<ImageAndTextCell> markedForDelete = new List<ImageAndTextCell> (); foreach (ImageAndTextCell cell in _refPool) { uint count = Messaging.UInt32_objc_msgSend (cell.Handle, selRetainCount); if (count == 1) markedForDelete.Add (cell); } foreach (ImageAndTextCell cell in markedForDelete) { _refPool.Remove (cell); cell.Dispose (); } } // Method 5 static IntPtr selCopyWithZone = Selector.GetHandle("copyWithZone:"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { IntPtr copyHandle = Messaging.IntPtr_objc_msgSendSuper_IntPtr(SuperHandle, selCopyWithZone, zone); ImageAndTextCell cell = new ImageAndTextCell(copyHandle) { Image = Image, }; _refPool.Add(cell); return cell; } Method 1: Increases the retain count of the unmanaged object. The unmanaged object will persist persist forever (I think? dealloc never called), and the managed object will be harvested early. Seems to be lose-lose all-around, but runs in practice. Method 2: Saves a reference of the managed object. The unmanaged object is left alone, and dealloc appears to be invoked at a reasonable time by the caller. At this point the managed object is released and disposed. This seems reasonable, but on the downside the base type's dealloc won't be run (I think?) Method 3: Increases the retain count and saves a reference. Unmanaged and managed objects leak forever. Method 4: Extends Method 3 by adding a cleanup function that is run periodically (e.g. during Init of each new ImageAndTextCell object). The cleanup function checks the retain counts of the stored objects. A retain count of 1 means the caller has released it, so we should as well. Should eliminate leaking in theory. Method 5: Attempt to invoke the copyWithZone method on the base type, and then construct a new ImageAndTextView object with the resulting handle. Seems to do the right thing (the base data is cloned). Internally, NSObject bumps the retain count on objects constructed like this, so we also use the PeriodicCleanup function to release these objects when they're no longer used. Based on the above, I believe Method 5 is the best approach since it should be the only one that results in a truly correct copy of the base type data, but I don't know if the approach is inherently dangerous (I am also making some assumptions about the underlying implementation of NSObject). So far nothing bad has happened "yet", but if anyone is able to vet my analysis then I would be more confident going forward.

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  • Database file is inexplicably locked during SQLite commit

    - by sweeney
    Hello, I'm performing a large number of INSERTS to a SQLite database. I'm using just one thread. I batch the writes to improve performance and have a bit of security in case of a crash. Basically I cache up a bunch of data in memory and then when I deem appropriate, I loop over all of that data and perform the INSERTS. The code for this is shown below: public void Commit() { using (SQLiteConnection conn = new SQLiteConnection(this.connString)) { conn.Open(); using (SQLiteTransaction trans = conn.BeginTransaction()) { using (SQLiteCommand command = conn.CreateCommand()) { command.CommandText = "INSERT OR IGNORE INTO [MY_TABLE] (col1, col2) VALUES (?,?)"; command.Parameters.Add(this.col1Param); command.Parameters.Add(this.col2Param); foreach (Data o in this.dataTemp) { this.col1Param.Value = o.Col1Prop; this. col2Param.Value = o.Col2Prop; command.ExecuteNonQuery(); } } this.TryHandleCommit(trans); } conn.Close(); } } I now employ the following gimmick to get the thing to eventually work: private void TryHandleCommit(SQLiteTransaction trans) { try { trans.Commit(); } catch (Exception e) { Console.WriteLine("Trying again..."); this.TryHandleCommit(trans); } } I create my DB like so: public DataBase(String path) { //build connection string SQLiteConnectionStringBuilder connString = new SQLiteConnectionStringBuilder(); connString.DataSource = path; connString.Version = 3; connString.DefaultTimeout = 5; connString.JournalMode = SQLiteJournalModeEnum.Persist; connString.UseUTF16Encoding = true; using (connection = new SQLiteConnection(connString.ToString())) { //check for existence of db FileInfo f = new FileInfo(path); if (!f.Exists) //build new blank db { SQLiteConnection.CreateFile(path); connection.Open(); using (SQLiteTransaction trans = connection.BeginTransaction()) { using (SQLiteCommand command = connection.CreateCommand()) { command.CommandText = DataBase.CREATE_MATCHES; command.ExecuteNonQuery(); command.CommandText = DataBase.CREATE_STRING_DATA; command.ExecuteNonQuery(); //TODO add logging } trans.Commit(); } connection.Close(); } } } I then export the connection string and use it to obtain new connections in different parts of the program. At seemingly random intervals, though at far too great a rate to ignore or otherwise workaround this problem, I get unhandled SQLiteException: Database file is locked. This occurs when I attempt to commit the transaction. No errors seem to occur prior to then. This does not always happen. Sometimes the whole thing runs without a hitch. No reads are being performed on these files before the commits finish. I have the very latest SQLite binary. I'm compiling for .NET 2.0. I'm using VS 2008. The db is a local file. All of this activity is encapsulated within one thread / process. Virus protection is off (though I think that was only relevant if you were connecting over a network?). As per Scotsman's post I have implemented the following changes: Journal Mode set to Persist DB files stored in C:\Docs + Settings\ApplicationData via System.Windows.Forms.Application.AppData windows call No inner exception Witnessed on two distinct machines (albeit very similar hardware and software) Have been running Process Monitor - no extraneous processes are attaching themselves to the DB files - the problem is definitely in my code... Does anyone have any idea whats going on here? I know I just dropped a whole mess of code, but I've been trying to figure this out for way too long. My thanks to anyone who makes it to the end of this question! brian UPDATES: Thanks for the suggestions so far! I've implemented many of the suggested changes. I feel that we are getting closer to the answer...however... The code above technically works however it is non-deterministic! It is not guaranteed to do anything aside from spin in neutral forever. In practice it seems to work somewhere between the 1st and 10th iteration. If i batch my commits at a reasonable interval damage will be mitigated but I really do not want to leave things in this state... More suggestions welcome!

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  • Trying to reduce the speed overhead of an almost-but-not-quite-int number class

    - by Fumiyo Eda
    I have implemented a C++ class which behaves very similarly to the standard int type. The difference is that it has an additional concept of "epsilon" which represents some tiny value that is much less than 1, but greater than 0. One way to think of it is as a very wide fixed point number with 32 MSBs (the integer parts), 32 LSBs (the epsilon parts) and a huge sea of zeros in between. The following class works, but introduces a ~2x speed penalty in the overall program. (The program includes code that has nothing to do with this class, so the actual speed penalty of this class is probably much greater than 2x.) I can't paste the code that is using this class, but I can say the following: +, -, +=, <, > and >= are the only heavily used operators. Use of setEpsilon() and getInt() is extremely rare. * is also rare, and does not even need to consider the epsilon values at all. Here is the class: #include <limits> struct int32Uepsilon { typedef int32Uepsilon Self; int32Uepsilon () { _value = 0; _eps = 0; } int32Uepsilon (const int &i) { _value = i; _eps = 0; } void setEpsilon() { _eps = 1; } Self operator+(const Self &rhs) const { Self result = *this; result._value += rhs._value; result._eps += rhs._eps; return result; } Self operator-(const Self &rhs) const { Self result = *this; result._value -= rhs._value; result._eps -= rhs._eps; return result; } Self operator-( ) const { Self result = *this; result._value = -result._value; result._eps = -result._eps; return result; } Self operator*(const Self &rhs) const { return this->getInt() * rhs.getInt(); } // XXX: discards epsilon bool operator<(const Self &rhs) const { return (_value < rhs._value) || (_value == rhs._value && _eps < rhs._eps); } bool operator>(const Self &rhs) const { return (_value > rhs._value) || (_value == rhs._value && _eps > rhs._eps); } bool operator>=(const Self &rhs) const { return (_value >= rhs._value) || (_value == rhs._value && _eps >= rhs._eps); } Self &operator+=(const Self &rhs) { this->_value += rhs._value; this->_eps += rhs._eps; return *this; } Self &operator-=(const Self &rhs) { this->_value -= rhs._value; this->_eps -= rhs._eps; return *this; } int getInt() const { return(_value); } private: int _value; int _eps; }; namespace std { template<> struct numeric_limits<int32Uepsilon> { static const bool is_signed = true; static int max() { return 2147483647; } } }; The code above works, but it is quite slow. Does anyone have any ideas on how to improve performance? There are a few hints/details I can give that might be helpful: 32 bits are definitely insufficient to hold both _value and _eps. In practice, up to 24 ~ 28 bits of _value are used and up to 20 bits of _eps are used. I could not measure a significant performance difference between using int32_t and int64_t, so memory overhead itself is probably not the problem here. Saturating addition/subtraction on _eps would be cool, but isn't really necessary. Note that the signs of _value and _eps are not necessarily the same! This broke my first attempt at speeding this class up. Inline assembly is no problem, so long as it works with GCC on a Core i7 system running Linux!

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  • C++0x rvalue references - lvalues-rvalue binding

    - by Doug
    This is a follow-on question to http://stackoverflow.com/questions/2748866/c0x-rvalue-references-and-temporaries In the previous question, I asked how this code should work: void f(const std::string &); //less efficient void f(std::string &&); //more efficient void g(const char * arg) { f(arg); } It seems that the move overload should probably be called because of the implicit temporary, and this happens in GCC but not MSVC (or the EDG front-end used in MSVC's Intellisense). What about this code? void f(std::string &&); //NB: No const string & overload supplied void g1(const char * arg) { f(arg); } void g2(const std::string & arg) { f(arg); } It seems that, based on the answers to my previous question that function g1 is legal (and is accepted by GCC 4.3-4.5, but not by MSVC). However, GCC and MSVC both reject g2 because of clause 13.3.3.1.4/3, which prohibits lvalues from binding to rvalue ref arguments. I understand the rationale behind this - it is explained in N2831 "Fixing a safety problem with rvalue references". I also think that GCC is probably implementing this clause as intended by the authors of that paper, because the original patch to GCC was written by one of the authors (Doug Gregor). However, I don't this is quite intuitive. To me, (a) a const string & is conceptually closer to a string && than a const char *, and (b) the compiler could create a temporary string in g2, as if it were written like this: void g2(const std::string & arg) { f(std::string(arg)); } Indeed, sometimes the copy constructor is considered to be an implicit conversion operator. Syntactically, this is suggested by the form of a copy constructor, and the standard even mentions this specifically in clause 13.3.3.1.2/4, where the copy constructor for derived-base conversions is given a higher conversion rank than other implicit conversions: A conversion of an expression of class type to the same class type is given Exact Match rank, and a conversion of an expression of class type to a base class of that type is given Conversion rank, in spite of the fact that a copy/move constructor (i.e., a user-defined conversion function) is called for those cases. (I assume this is used when passing a derived class to a function like void h(Base), which takes a base class by value.) Motivation My motivation for asking this is something like the question asked in http://stackoverflow.com/questions/2696156/how-to-reduce-redundant-code-when-adding-new-c0x-rvalue-reference-operator-over ("How to reduce redundant code when adding new c++0x rvalue reference operator overloads"). If you have a function that accepts a number of potentially-moveable arguments, and would move them if it can (e.g. a factory function/constructor: Object create_object(string, vector<string>, string) or the like), and want to move or copy each argument as appropriate, you quickly start writing a lot of code. If the argument types are movable, then one could just write one version that accepts the arguments by value, as above. But if the arguments are (legacy) non-movable-but-swappable classes a la C++03, and you can't change them, then writing rvalue reference overloads is more efficient. So if lvalues did bind to rvalues via an implicit copy, then you could write just one overload like create_object(legacy_string &&, legacy_vector<legacy_string> &&, legacy_string &&) and it would more or less work like providing all the combinations of rvalue/lvalue reference overloads - actual arguments that were lvalues would get copied and then bound to the arguments, actual arguments that were rvalues would get directly bound. Questions My questions are then: Is this a valid interpretation of the standard? It seems that it's not the conventional or intended one, at any rate. Does it make intuitive sense? Is there a problem with this idea that I"m not seeing? It seems like you could get copies being quietly created when that's not exactly expected, but that's the status quo in places in C++03 anyway. Also, it would make some overloads viable when they're currently not, but I don't see it being a problem in practice. Is this a significant enough improvement that it would be worth making e.g. an experimental patch for GCC?

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  • jQuery programming style?

    - by Sam Dufel
    I was recently asked to fix something on a site which I haven't worked on before. I haven't really worked with jQuery that much, but I figured I'd take a look and see if I could fix it. I've managed to mostly clear up the problem, but I'm still horrified at the way they chose to build this site. On document load, they replace the click() method of every anchor tag and form element with the same massive function. When clicked, that function then checks if the tag has one of a few different attributes (non-standard attributes, even), and does a variety of different tasks depending on what attributes exist and what their values are. Some hyperlinks have an attribute on them called 'ajaxrel', which makes the click() function look for another (hidden) hyperlink with an ID specified by the ajaxrel attribute, and then calls the click() function for that other hyperlink (which was also modified by this same click() function). On the server side, all the php files are quite long and have absolutely no indentation. This whole site has been a nightmare to debug. Is this standard jQuery practice? This navigation scheme seems terrible. Does anyone else actually use jQuery this way? I'd like to start incorporating it into my projects, but looking at this site is giving me a serious headache. Here's the click() function for hyperlinks: function ajaxBoxA(theElement, urltosend, ajaxbox, dialogbox) { if ($(theElement).attr("href") != undefined) var urltosend = $(theElement).attr("href"); if ($(theElement).attr('toajaxbox') != undefined) var ajaxbox = $(theElement).attr('toajaxbox'); // check to see if dialog box is called for. if ($(theElement).attr('dialogbox') != undefined) var dialogbox = $(theElement).attr('dialogbox'); var dodialog = 0; if (dialogbox != undefined) { // if dialogbox doesn't exist, then flag to create dialog box. var isDiaOpen = $('[ajaxbox="' + ajaxbox + '"]').parent().parent().is(".ui-dialog-container"); dodialog = 1; if (isDiaOpen) { dodialog = 0; } dialogbox = parseUri(dialogbox); dialogoptions = { close: function () { // $("[id^=hierarchy]",this).NestedSortableDestroy(); $(this).dialog('destroy').remove() } }; for ( var keyVar in dialogbox['queryKey'] ) eval( "dialogoptions." + keyVar + " = dialogbox['queryKey'][keyVar]"); }; $("body").append("<div id='TB_load'><img src='"+imgLoader.src+"' /></div>"); $('#TB_load').show(); if (urltosend.search(/\?/) > 0) { urltosend = urltosend + "&-ajax=1"; } else { urltosend = urltosend + "?-ajax=1"; } if ($('[ajaxbox="' + ajaxbox + '"]').length) { $('[ajaxbox="' + ajaxbox + '"]').each( function () { $(this).empty(); }); }; $.ajax({ type: "GET", url: urltosend, data: "", async: false, dataType: "html", success: function (html) { var re = /^<toajaxbox>(.*?)<\/toajaxbox>+(.*)/; if (re.test(html)) { var match = re.exec(html); ajaxbox = match[1]; html = Right(html, String(html).length - String(match[1]).length); } var re = /^<header>(.*?)<\/header>+(.*)/; if (re.test(html)) { var match = re.exec(html); window.location = match[1]; return false; } if (html.length > 0) { var newHtml = $(html); if ($('[ajaxbox="' + ajaxbox + '"]').length) { $('[ajaxbox="' + ajaxbox + '"]').each( function () { $(this).replaceWith(newHtml).ready( function () { ajaxBoxInit(newHtml) if (window.ajaxboxsuccess) ajaxboxsuccess(newHtml); }); }); if ($('[ajaxdialog="' + ajaxbox + '"]').length = 0) { if (dodialog) $(newHtml).wrap("<div class='flora ui-dialog-content' ajaxdialog='" + ajaxbox + "' style='overflow:auto;'></div>").parent().dialog(dialogoptions); } } else { $("body").append(newHtml).ready( function () { ajaxBoxInit(newHtml); if (window.ajaxboxsuccess) ajaxboxsuccess(newHtml); }); if (dodialog) $(newHtml).wrap("<div class='flora ui-dialog-content' ajaxdialog='" + ajaxbox + "' style='overflow:auto;'></div>").parent().dialog(dialogoptions); } } var rel = $(theElement).attr('ajaxtriggerrel'); if (rel != undefined) $('a[ajaxrel="' + rel + '"]').click(); tb_remove(); return false; }, complete: function () { $("#TB_load").remove(); } }); return false; }

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  • java/Swing issue with paintComponent

    - by user310254
    The issue I'm having is issue with is I'm trying to get the paintComponent to draw the circle only when the mouse is clicked, dragged, then let go. However inside my paintPanel class I have to initialize the object I've created (ex. movedCircle myCircle = new movedCircle(0,0,0,0);) just creating the object movedCircle myCircle; gives an error until I actually fully initialize the object with a value. What I'm looking for: What's considered the best practice for this issue. I don't want to draw anything unnecessary before it is needed. The way I know how to fix it: boolean values inside of paintComponent so that way it doesn't draw until somethings actually there. import javax.swing.*; import java.awt.*; import java.awt.event.*; public class drawCircle extends JFrame{ private JPanel myPanel = new paintPanel(); public drawCircle(){ add(myPanel); } private class paintPanel extends JPanel{ private int x1, y1, x2, y2; movedText myText = new movedText(0,0,0,0); movedCircle myCircle = new movedCircle(0,0,0,0); public paintPanel(){ addMouseListener(new MouseAdapter(){ public void mousePressed(MouseEvent e){ x1 = e.getX(); y1 = e.getY(); myCircle = new movedCircle(x1, y1, 0, 0); repaint(); } public void mouseReleased(MouseEvent e){ x2 = e.getX(); y2 = e.getY(); myCircle = new movedCircle(x1, y1, x2, y2); repaint(); } }); addMouseMotionListener(new MouseMotionAdapter(){ public void mouseDragged(MouseEvent e){ x2 = e.getX(); y2 = e.getY(); myText = new movedText(x1, y1, x2, y2); myCircle = new movedCircle(x1, y1, x2, y2); repaint(); } public void mouseMoved(MouseEvent e){ x1 = e.getX(); y1 = e.getY(); x2 = 0; y2 = 0; myText = new movedText(x1, y1, x2, y2); repaint(); } }); } protected void paintComponent(Graphics g){ super.paintComponent(g); //draw oval after mouse released myText.paintText(g); myCircle.paintCircle(g); } } class movedCircle{ private int x1, y1, x2, y2; public movedCircle(int x1, int y1, int x2, int y2){ this.x1 = x1; this.y1 = y1; this.x2 = x2; this.y2 = y2; } public void paintCircle(Graphics g){ g.drawOval(x1, y1, x2 - x1, y2 - y1); } } class movedText{ private int x1, y1, x2, y2; public movedText(int x1, int y1, int x2, int y2){ this.x1 = x1; this.y1 = y1; this.x2 = x2; this.y2 = y2; } public void paintText(Graphics g){ g.drawString("x1: "+x1+" y1: "+y1+" x2: "+x2+" y2: "+y2, x1, y1); } } class RedSquare{ private int xPos = 50; private int yPos = 50; private int width = 20; private int height = 20; public void setX(int xPos){ this.xPos = xPos; } public int getX(){ return xPos; } public void setY(int yPos){ this.yPos = yPos; } public int getY(){ return yPos; } public int getWidth(){ return width; } public int getHeight(){ return height; } public void paintSquare(Graphics g){ g.setColor(Color.RED); g.fillRect(xPos,yPos,width,height); g.setColor(Color.BLACK); g.drawRect(xPos,yPos,width,height); } } public static void main(String[] args){ JFrame frame = new drawCircle(); frame.setTitle("Is in ellipse? Demo"); frame.setSize(400, 400); frame.setLocationRelativeTo(null); frame.setDefaultCloseOperation(JFrame.EXIT_ON_CLOSE); frame.setVisible(true); } }

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  • Java RMI cannot connect to host from external client.

    - by Koe
    I've been using RMI in this project for a while. I've gotten the client program to connect (amongst other things) to the server when running it over my LAN, however when running it over the internet I'm running into the following exception: java.rmi.ConnectException: Connection refused to host: (private IP of host machine); nested exception is: java.net.ConnectException: Connection timed out: connect at sun.rmi.transport.tcp.TCPEndpoint.newSocket(Unknown Source) at sun.rmi.transport.tcp.TCPChannel.createConnection(Unknown Source) at sun.rmi.transport.tcp.TCPChannel.newConnection(Unknown Source) at sun.rmi.server.UnicastRef.invoke(Unknown Source) at java.rmi.server.RemoteObjectInvocationHandler.invokeRemoteMethod(Unknown Source) at java.rmi.server.RemoteObjectInvocationHandler.invoke(Unknown Source) at $Proxy1.ping(Unknown Source) at client.Launcher$PingLabel.runPing(Launcher.java:366) at client.Launcher$PingLabel.<init>(Launcher.java:353) at client.Launcher.setupContentPane(Launcher.java:112) at client.Launcher.<init>(Launcher.java:99) at client.Launcher.main(Launcher.java:59) Caused by: java.net.ConnectException: Connection timed out: connect at java.net.PlainSocketImpl.socketConnect(Native Method) at java.net.PlainSocketImpl.doConnect(Unknown Source) at java.net.PlainSocketImpl.connectToAddress(Unknown Source) at java.net.PlainSocketImpl.connect(Unknown Source) at java.net.SocksSocketImpl.connect(Unknown Source) at java.net.Socket.connect(Unknown Source) at java.net.Socket.connect(Unknown Source) at java.net.Socket.<init>(Unknown Source) at java.net.Socket.<init>(Unknown Source) at sun.rmi.transport.proxy.RMIDirectSocketFactory.createSocket(Unknown Source) at sun.rmi.transport.proxy.RMIMasterSocketFactory.createSocket(Unknown Source) ... 12 more This error is remeniscent of my early implementation of RMI and I can obtain the error verbatum if I run the client locally without the server program running as well. To me Connection Timed Out means a problem with the server's response. Here's the client initiation: public static void main(String[] args) { try { String host = "<WAN IP>"; Registry registry = LocateRegistry.getRegistry(host, 1099); Login lstub = (Login) registry.lookup("Login Server"); Information istub = (Information) registry.lookup("Game Server"); new Launcher(istub, lstub); } catch (RemoteException e) { System.err.println("Client exception: " + e.toString()); e.printStackTrace(); } catch (NotBoundException e) { System.err.println("Client exception: " + e.toString()); e.printStackTrace(); } } Interestingly enough no Remote Exception is thrown here. Here's the server initiation: public static void main(String args[]) { try { GameServer gobj = new GameServer(); Information gstub = (Information) UnicastRemoteObject.exportObject( gobj, 1099); Registry registry = LocateRegistry.createRegistry(1099); registry.bind("Game Server", gstub); LoginServer lobj = new LoginServer(gobj); Login lstub = (Login) UnicastRemoteObject.exportObject(lobj, 7099); // Bind the remote object's stub in the registry registry.bind("Login Server", lstub); System.out.println("Server ready"); } catch (Exception e) { System.err.println("Server exception: " + e.toString()); e.printStackTrace(); } } Bad practice with the catch(Exception e) I know but bear with me. Up to this stage I know it works fine over the LAN, here's where the exception occurs over the WAN and is the first place a method in the server is called: private class PingLabel extends JLabel { private static final long serialVersionUID = 1L; public PingLabel() { super(""); runPing(); } public void setText(String text) { super.setText("Ping: " + text + "ms"); } public void runPing() { try { PingThread pt = new PingThread(); gameServer.ping(); pt.setRecieved(true); setText("" + pt.getTime()); } catch (RemoteException e) { e.printStackTrace(); } } } That's a label placed on the launcher as a ping test. the method ping(), in gameserver does nothing, as in is a null method. It's worth noting also that ports 1099 and 7099 are forwarded to the server machine (which should be obvious from the stack trace). Can anyone see anyting I'm missing/doing wrong? If you need any more information just ask. EDIT: I'm practically certain the problem has nothing to do with my router settings. When disabling my port forwarding settings I get a slightly different error: Client exception: java.rmi.ConnectException: Connection refused to host: (-WAN IP NOT LOCAL IP-); but it appears both on the machine locally connected to the server and on the remote machine. In addition, I got it to work seamlessly when connecting the server straight tho the modem (cutting out the router. I can only conclude the problem is in my router's settings but can't see where (I've checked and double checked the port forwarding page). That's the only answer i can come up with.

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  • Object oriented n-tier design. Am I abstracting too much? Or not enough?

    - by max
    Hi guys, I'm building my first enterprise grade solution (at least I'm attempting to make it enterprise grade). I'm trying to follow best practice design patterns but am starting to worry that I might be going too far with abstraction. I'm trying to build my asp.net webforms (in C#) app as an n-tier application. I've created a Data Access Layer using an XSD strongly-typed dataset that interfaces with a SQL server backend. I access the DAL through some Business Layer Objects that I've created on a 1:1 basis to the datatables in the dataset (eg, a UsersBLL class for the Users datatable in the dataset). I'm doing checks inside the BLL to make sure that data passed to DAL is following the business rules of the application. That's all well and good. Where I'm getting stuck though is the point at which I connect the BLL to the presentation layer. For example, my UsersBLL class deals mostly with whole datatables, as it's interfacing with the DAL. Should I now create a separate "User" (Singular) class that maps out the properties of a single user, rather than multiple users? This way I don't have to do any searching through datatables in the presentation layer, as I could use the properties created in the User class. Or would it be better to somehow try to handle this inside the UsersBLL? Sorry if this sounds a little complicated... Below is the code from the UsersBLL: using System; using System.Data; using PedChallenge.DAL.PedDataSetTableAdapters; [System.ComponentModel.DataObject] public class UsersBLL { private UsersTableAdapter _UsersAdapter = null; protected UsersTableAdapter Adapter { get { if (_UsersAdapter == null) _UsersAdapter = new UsersTableAdapter(); return _UsersAdapter; } } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Select, true)] public PedChallenge.DAL.PedDataSet.UsersDataTable GetUsers() { return Adapter.GetUsers(); } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Select, false)] public PedChallenge.DAL.PedDataSet.UsersDataTable GetUserByUserID(int userID) { return Adapter.GetUserByUserID(userID); } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Select, false)] public PedChallenge.DAL.PedDataSet.UsersDataTable GetUsersByTeamID(int teamID) { return Adapter.GetUsersByTeamID(teamID); } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Select, false)] public PedChallenge.DAL.PedDataSet.UsersDataTable GetUsersByEmail(string Email) { return Adapter.GetUserByEmail(Email); } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Insert, true)] public bool AddUser(int? teamID, string FirstName, string LastName, string Email, string Role, int LocationID) { // Create a new UsersRow instance PedChallenge.DAL.PedDataSet.UsersDataTable Users = new PedChallenge.DAL.PedDataSet.UsersDataTable(); PedChallenge.DAL.PedDataSet.UsersRow user = Users.NewUsersRow(); if (UserExists(Users, Email) == true) return false; if (teamID == null) user.SetTeamIDNull(); else user.TeamID = teamID.Value; user.FirstName = FirstName; user.LastName = LastName; user.Email = Email; user.Role = Role; user.LocationID = LocationID; // Add the new user Users.AddUsersRow(user); int rowsAffected = Adapter.Update(Users); // Return true if precisely one row was inserted, // otherwise false return rowsAffected == 1; } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Update, true)] public bool UpdateUser(int userID, int? teamID, string FirstName, string LastName, string Email, string Role, int LocationID) { PedChallenge.DAL.PedDataSet.UsersDataTable Users = Adapter.GetUserByUserID(userID); if (Users.Count == 0) // no matching record found, return false return false; PedChallenge.DAL.PedDataSet.UsersRow user = Users[0]; if (teamID == null) user.SetTeamIDNull(); else user.TeamID = teamID.Value; user.FirstName = FirstName; user.LastName = LastName; user.Email = Email; user.Role = Role; user.LocationID = LocationID; // Update the product record int rowsAffected = Adapter.Update(user); // Return true if precisely one row was updated, // otherwise false return rowsAffected == 1; } [System.ComponentModel.DataObjectMethodAttribute (System.ComponentModel.DataObjectMethodType.Delete, true)] public bool DeleteUser(int userID) { int rowsAffected = Adapter.Delete(userID); // Return true if precisely one row was deleted, // otherwise false return rowsAffected == 1; } private bool UserExists(PedChallenge.DAL.PedDataSet.UsersDataTable users, string email) { // Check if user email already exists foreach (PedChallenge.DAL.PedDataSet.UsersRow userRow in users) { if (userRow.Email == email) return true; } return false; } } Some guidance in the right direction would be greatly appreciated!! Thanks all! Max

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  • C++ Program performs better when piped

    - by ET1 Nerd
    I haven't done any programming in a decade. I wanted to get back into it, so I made this little pointless program as practice. The easiest way to describe what it does is with output of my --help codeblock: ./prng_bench --help ./prng_bench: usage: ./prng_bench $N $B [$T] This program will generate an N digit base(B) random number until all N digits are the same. Once a repeating N digit base(B) number is found, the following statistics are displayed: -Decimal value of all N digits. -Time & number of tries taken to randomly find. Optionally, this process is repeated T times. When running multiple repititions, averages for all N digit base(B) numbers are displayed at the end, as well as total time and total tries. My "problem" is that when the problem is "easy", say a 3 digit base 10 number, and I have it do a large number of passes the "total time" is less when piped to grep. ie: command ; command |grep took : ./prng_bench 3 10 999999 ; ./prng_bench 3 10 999999|grep took .... Pass# 999999: All 3 base(10) digits = 3 base(10). Time: 0.00005 secs. Tries: 23 It took 191.86701 secs & 99947208 tries to find 999999 repeating 3 digit base(10) numbers. An average of 0.00019 secs & 99 tries was needed to find each one. It took 159.32355 secs & 99947208 tries to find 999999 repeating 3 digit base(10) numbers. If I run the same command many times w/o grep time is always VERY close. I'm using srand(1234) for now, to test. The code between my calls to clock_gettime() for start and stop do not involve any stream manipulation, which would obviously affect time. I realize this is an exercise in futility, but I'd like to know why it behaves this way. Below is heart of the program. Here's a link to the full source on DB if anybody wants to compile and test. https://www.dropbox.com/s/6olqnnjf3unkm2m/prng_bench.cpp clock_gettime() requires -lrt. for (int pass_num=1; pass_num<=passes; pass_num++) { //Executes $passes # of times. clock_gettime(CLOCK_PROCESS_CPUTIME_ID, &temp_time); //get time start_time = timetodouble(temp_time); //convert time to double, store as start_time for(i=1, tries=0; i!=0; tries++) { //loops until 'comparison for' fully completes. counts reps as 'tries'. <------------ for (i=0; i<Ndigits; i++) //Move forward through array. | results[i]=(rand()%base); //assign random num of base to element (digit). | /*for (i=0; i<Ndigits; i++) //---Debug Lines--------------- | std::cout<<" "<<results[i]; //---a LOT of output.---------- | std::cout << "\n"; //---Comment/decoment to disable/enable.*/ // | for (i=Ndigits-1; i>0 && results[i]==results[0]; i--); //Move through array, != element breaks & i!=0, new digits drawn. -| } //If all are equal i will be 0, nested for condition satisfied. -| clock_gettime(CLOCK_PROCESS_CPUTIME_ID, &temp_time); //get time draw_time = (timetodouble(temp_time) - start_time); //convert time to dbl, subtract start_time, set draw_time to diff. total_time += draw_time; //add time for this pass to total. total_tries += tries; //add tries for this pass to total. /*Formated output for each pass: Pass# ---: All -- base(--) digits = -- base(10) Time: ----.---- secs. Tries: ----- (LINE) */ std::cout<<"Pass# "<<std::setw(width_pass)<<pass_num<<": All "<<Ndigits<<" base("<<base<<") digits = " <<std::setw(width_base)<<results[0]<<" base(10). Time: "<<std::setw(width_time)<<draw_time <<" secs. Tries: "<<tries<<"\n"; } if(passes==1) return 0; //No need for totals and averages of 1 pass. /* It took ----.---- secs & ------ tries to find --- repeating -- digit base(--) numbers. (LINE) An average of ---.---- secs & ---- tries was needed to find each one. (LINE)(LINE) */ std::cout<<"It took "<<total_time<<" secs & "<<total_tries<<" tries to find " <<passes<<" repeating "<<Ndigits<<" digit base("<<base<<") numbers.\n" <<"An average of "<<total_time/passes<<" secs & "<<total_tries/passes <<" tries was needed to find each one. \n\n"; return 0;

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  • Accessing local variable doesn't improve performance

    - by NicMagnier
    The short version Why is this code: var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; More performant than this one? var index = Math.floor(ref_index) * 4; The long version This week, the author of Impact js published an article about some rendering issue: http://www.phoboslab.org/log/2012/09/drawing-pixels-is-hard In the article there was the source of a function to scale an image by accessing pixels in the canvas. I wanted to suggest some traditional ways to optimize this kind of code so that the scaling would be shorter at loading time. But after testing it my result was most of the time worst that the original function. Guessing this was the JavaScript engine that was doing some smart optimization I tried to understand a bit more what was going on so I did a bunch of test. But my results are quite confusing and I would need some help to understand what's going on. I have a test page here: http://www.mx981.com/stuff/resize_bench/test.html jsPerf: http://jsperf.com/local-variable-due-to-the-scope-lookup To start the test, click the picture and the results will appear in the console. There are three different versions: The original code: for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; var indexScaled = (y * widthScaled + x) * 4; scaledPixels.data[ indexScaled ] = origPixels.data[ index ]; scaledPixels.data[ indexScaled+1 ] = origPixels.data[ index+1 ]; scaledPixels.data[ indexScaled+2 ] = origPixels.data[ index+2 ]; scaledPixels.data[ indexScaled+3 ] = origPixels.data[ index+3 ]; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance One of my attempt to optimize it: var ref_index = 0; var ref_indexScaled = 0 var ref_step = 1 / scale; for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = Math.floor(ref_index) * 4; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+1 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+2 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+3 ]; ref_index+= ref_step; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance The same optimized code but with recalculating the index variable each time (Hybrid) var ref_index = 0; var ref_indexScaled = 0 var ref_step = 1 / scale; for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+1 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+2 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+3 ]; ref_index+= ref_step; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance The only difference in the two last one is the calculation of the 'index' variable. And to my surprise the optimized version is slower in most browsers (except opera). Results of personal testing (not the jsPerf tests): Opera Original: 8668ms Optimized: 932ms Hybrid: 8696ms Chrome Original: 139ms Optimized: 145ms Hybrid: 136ms Safari Original: 433ms Optimized: 853ms Hybrid: 451ms Firefox Original: 343ms Optimized: 422ms Hybrid: 350ms After digging around, it seems an usual good practice is to access mainly local variable due to the scope lookup. Because The optimized version only call one local variable it should be faster that the Hybrid code which call multiple variable and object in addition to the various operation involved. So why the "optimized" version is slower? I thought that it might be because some JavaScript engine don't optimize the Optimized version because it is not hot enough but after using --trace-opt in chrome, it seems all version are properly compiled by V8. At this point I am a bit clueless and wonder if somebody would know what is going on? I did also some more test cases in this page: http://www.mx981.com/stuff/resize_bench/index.html

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  • ERR_INCOMPLETE_CHUNKED_ENCODING apache 2.4

    - by Bujanca Mihai
    I upgraded my Ubuntu server to 14.04 and Apache 2.4.7. Now my images don't load and console yields net::ERR_INCOMPLETE_CHUNKED_ENCODING. Also, I can sometimes see some of the images load for a little while (1 sec max) and then they disappear. .htaccess RewriteEngine On # Serve the favicon file from img folder RewriteCond %{REQUEST_URI} ^/favicon.ico$ RewriteRule ^(.*)$ /img/$1 [NC,L] # Redirect HTTP traffic to WWW subdomain RewriteCond %{HTTPS} off [NC] RewriteCond %{HTTP_HOST} !^www\. [NC] RewriteRule ^(.*)$ http://www.%{HTTP_HOST}/$1 [R=301,L] # Redirect HTTPS traffic to WWW subdomain RewriteCond %{HTTPS} on [NC] RewriteCond %{HTTP_HOST} !^www\. [NC] RewriteRule ^(.*)$ https://www.%{HTTP_HOST}/$1 [R=301,L] # Auto Versioning rules RewriteCond %{REQUEST_FILENAME} !-s RewriteRule ^(.*)\.[\d]+\.(css|js)$ $1.$2 [L] # Default Zend rewrite rules RewriteCond %{REQUEST_FILENAME} -s [OR] RewriteCond %{REQUEST_FILENAME} -l [OR] RewriteCond %{REQUEST_FILENAME} -d RewriteRule ^.*$ - [NC,L] RewriteRule ^.*$ index.php [NC,L] VHost <VirtualHost *:80> ServerAdmin admin@localhost ServerName localhost DocumentRoot /home/mihai/ARTD/www/public/website # Omit this in production environment SetEnv APPLICATION_ENV local <Directory /home/mihai/ARTD/www/public/website > Options Indexes FollowSymLinks MultiViews AllowOverride All #Order deny,allow #Allow from all Require all granted </Directory> <IfModule mod_php5.c> php_value memory_limit 128M php_value upload_max_filesize 20M php_value post_max_size 20M </IfModule> ErrorLog /var/log/apache2/ARTD-error.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn CustomLog /var/log/apache2/ARTD-access.log combined </VirtualHost> <IfModule mod_ssl.c> <VirtualHost *:443> ServerAdmin admin@localhost ServerName localhost DocumentRoot /home/mihai/ARTD/www/public/website # Omit this in production environment SetEnv APPLICATION_ENV local <Directory /home/mihai/ARTD/www/public/website > Options Indexes FollowSymLinks MultiViews AllowOverride All #Order deny,allow #Allow from all Require all granted </Directory> <IfModule mod_php5.c> php_value memory_limit 128M php_value upload_max_filesize 20M php_value post_max_size 20M </IfModule> ErrorLog /var/log/apache2/ARTD-ssl-error.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn CustomLog /var/log/apache2/ARTD.log combined # SSL Engine Switch: # Enable/Disable SSL for this virtual host. SSLEngine on # A self-signed (snakeoil) certificate can be created by installing # the ssl-cert package. See # /usr/share/doc/apache2.2-common/README.Debian.gz for more info. # If both key and certificate are stored in the same file, only the # SSLCertificateFile directive is needed. SSLCertificateFile /etc/ssl/certs/ssl-cert-snakeoil.pem SSLCertificateKeyFile /etc/ssl/private/ssl-cert-snakeoil.key # Server Certificate Chain: # Point SSLCertificateChainFile at a file containing the # concatenation of PEM encoded CA certificates which form the # certificate chain for the server certificate. Alternatively # the referenced file can be the same as SSLCertificateFile # when the CA certificates are directly appended to the server # certificate for convinience. #SSLCertificateChainFile /etc/apache2/ssl.crt/server-ca.crt # Certificate Authority (CA): # Set the CA certificate verification path where to find CA # certificates for client authentication or alternatively one # huge file containing all of them (file must be PEM encoded) # Note: Inside SSLCACertificatePath you need hash symlinks # to point to the certificate files. Use the provided # Makefile to update the hash symlinks after changes. #SSLCACertificatePath /etc/ssl/certs/ #SSLCACertificateFile /etc/apache2/ssl.crt/ca-bundle.crt # Certificate Revocation Lists (CRL): # Set the CA revocation path where to find CA CRLs for client # authentication or alternatively one huge file containing all # of them (file must be PEM encoded) # Note: Inside SSLCARevocationPath you need hash symlinks # to point to the certificate files. Use the provided # Makefile to update the hash symlinks after changes. #SSLCARevocationPath /etc/apache2/ssl.crl/ #SSLCARevocationFile /etc/apache2/ssl.crl/ca-bundle.crl # Client Authentication (Type): # Client certificate verification type and depth. Types are # none, optional, require and optional_no_ca. Depth is a # number which specifies how deeply to verify the certificate # issuer chain before deciding the certificate is not valid. #SSLVerifyClient require #SSLVerifyDepth 10 # Access Control: # With SSLRequire you can do per-directory access control based # on arbitrary complex boolean expressions containing server # variable checks and other lookup directives. The syntax is a # mixture between C and Perl. See the mod_ssl documentation # for more details. #<Location /> #SSLRequire ( %{SSL_CIPHER} !~ m/^(EXP|NULL)/ \ # and %{SSL_CLIENT_S_DN_O} eq "Snake Oil, Ltd." \ # and %{SSL_CLIENT_S_DN_OU} in {"Staff", "CA", "Dev"} \ # and %{TIME_WDAY} >= 1 and %{TIME_WDAY} <= 5 \ # and %{TIME_HOUR} >= 8 and %{TIME_HOUR} <= 20 ) \ # or %{REMOTE_ADDR} =~ m/^192\.76\.162\.[0-9]+$/ #</Location> # SSL Engine Options: # Set various options for the SSL engine. # o FakeBasicAuth: # Translate the client X.509 into a Basic Authorisation. This means that # the standard Auth/DBMAuth methods can be used for access control. The # user name is the `one line' version of the client's X.509 certificate. # Note that no password is obtained from the user. Every entry in the user # file needs this password: `xxj31ZMTZzkVA'. # o ExportCertData: # This exports two additional environment variables: SSL_CLIENT_CERT and # SSL_SERVER_CERT. These contain the PEM-encoded certificates of the # server (always existing) and the client (only existing when client # authentication is used). This can be used to import the certificates # into CGI scripts. # o StdEnvVars: # This exports the standard SSL/TLS related `SSL_*' environment variables. # Per default this exportation is switched off for performance reasons, # because the extraction step is an expensive operation and is usually # useless for serving static content. So one usually enables the # exportation for CGI and SSI requests only. # o StrictRequire: # This denies access when "SSLRequireSSL" or "SSLRequire" applied even # under a "Satisfy any" situation, i.e. when it applies access is denied # and no other module can change it. # o OptRenegotiate: # This enables optimized SSL connection renegotiation handling when SSL # directives are used in per-directory context. #SSLOptions +FakeBasicAuth +ExportCertData +StrictRequire #<FilesMatch "\.(cgi|shtml|phtml|php)$"> # SSLOptions +StdEnvVars #</FilesMatch> # SSL Protocol Adjustments: # The safe and default but still SSL/TLS standard compliant shutdown # approach is that mod_ssl sends the close notify alert but doesn't wait for # the close notify alert from client. When you need a different shutdown # approach you can use one of the following variables: # o ssl-unclean-shutdown: # This forces an unclean shutdown when the connection is closed, i.e. no # SSL close notify alert is send or allowed to received. This violates # the SSL/TLS standard but is needed for some brain-dead browsers. Use # this when you receive I/O errors because of the standard approach where # mod_ssl sends the close notify alert. # o ssl-accurate-shutdown: # This forces an accurate shutdown when the connection is closed, i.e. a # SSL close notify alert is send and mod_ssl waits for the close notify # alert of the client. This is 100% SSL/TLS standard compliant, but in # practice often causes hanging connections with brain-dead browsers. Use # this only for browsers where you know that their SSL implementation # works correctly. # Notice: Most problems of broken clients are also related to the HTTP # keep-alive facility, so you usually additionally want to disable # keep-alive for those clients, too. Use variable "nokeepalive" for this. # Similarly, one has to force some clients to use HTTP/1.0 to workaround # their broken HTTP/1.1 implementation. Use variables "downgrade-1.0" and # "force-response-1.0" for this. #BrowserMatch ".*MSIE.*" \ # nokeepalive ssl-unclean-shutdown \ # downgrade-1.0 force-response-1.0 </VirtualHost> </IfModule> logs Apache/2.4.7 (Ubuntu) PHP/5.5.9-1ubuntu4.3 OpenSSL/1.0.1f (internal dummy connection) 127.0.0.1 - - [25/Aug/2014:13:09:53 +0300] "GET /img/header/top-nav-separator.png HTTP/1.1" 200 462 "https://localhost/art" "Mozilla/5.0 (X11; Linux x86_64) AppleWebKit/537.36 (KHTML, like Gecko) Chrome/34.0.1847.132 Safari/537.36"

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  • IIS Strategies for Accessing Secured Network Resources

    - by ErikE
    Problem: A user connects to a service on a machine, such as an IIS web site or a SQL Server database. The site or the database need to gain access to network resources such as file shares (the most common) or a database on a different server. Permission is denied. This is because the user the service is running under doesn't have network permissions in the first place, or if it does, it doesn't have rights to access the remote resource. I keep running into this problem over and over again and am tired of not having a really solid way of handling it. Here are some workarounds I'm aware of: Run IIS as a custom-created domain user who is granted high permissions If permissions are granted one file share at a time, then every time I want to read from a new share, I would have to ask a network admin to add it for me. Eventually, with many web sites reading from many shares, it is going to get really complicated. If permissions are just opened up wide for the user to access any file shares in our domain, then this seems like an unnecessary security surface area to present. This also applies to all the sites running on IIS, rather than just the selected site or virtual directory that needs the access, a further surface area problem. Still use the IUSR account but give it network permissions and set up the same user name on the remote resource (not a domain user, a local user) This also has its problems. For example, there's a file share I am using that I have full rights to for sharing, but I can't log in to the machine. So I have to find the right admin and ask him to do it for me. Any time something has to change, it's another request to an admin. Allow IIS users to connect as anonymous, but set the account used for anonymous access to a high-privilege one This is even worse than giving the IIS IUSR full privileges, because it means my web site can't use any kind of security in the first place. Connect using Kerberos, then delegate This sounds good in principle but has all sorts of problems. First of all, if you're using virtual web sites where the domain name you connect to the site with is not the base machine name (as we do frequently), then you have to set up a Service Principal Name on the webserver using Microsoft's SetSPN utility. It's complicated and apparently prone to errors. Also, you have to ask your network/domain admin to change security policy for both the web server and the domain account so they are "trusted for delegation." If you don't get everything perfectly right, suddenly your intended Kerberos authentication is NTLM instead, and you can only impersonate rather than delegate, and thus no reaching out over the network as the user. Also, this method can be problematic because sometimes you need the web site or database to have permissions that the connecting user doesn't have. Create a service or COM+ application that fetches the resource for the web site Services and COM+ packages are run with their own set of credentials. Running as a high-privilege user is okay since they can do their own security and deny requests that are not legitimate, putting control in the hands of the application developer instead of the network admin. Problems: I am using a COM+ package that does exactly this on Windows Server 2000 to deliver highly sensitive images to a secured web application. I tried moving the web site to Windows Server 2003 and was suddenly denied permission to instantiate the COM+ object, very likely registry permissions. I trolled around quite a bit and did not solve the problem, partly because I was reluctant to give the IUSR account full registry permissions. That seems like the same bad practice as just running IIS as a high-privilege user. Note: This is actually really simple. In a programming language of your choice, you create a class with a function that returns an instance of the object you want (an ADODB.Connection, for example), and build a dll, which you register as a COM+ object. In your web server-side code, you create an instance of the class and use the function, and since it is running under a different security context, calls to network resources work. Map drive letters to shares This could theoretically work, but in my mind it's not really a good long-term strategy. Even though mappings can be created with specific credentials, and this can be done by others than a network admin, this also is going to mean that there are either way too many shared drives (small granularity) or too much permission is granted to entire file servers (large granularity). Also, I haven't figured out how to map a drive so that the IUSR gets the drives. Mapping a drive is for the current user, I don't know the IUSR account password to log in as it and create the mappings. Move the resources local to the web server/database There are times when I've done this, especially with Access databases. Does the database have to live out on the file share? Sometimes, it was just easiest to move the database to the web server or to the SQL database server (so the linked server to it would work). But I don't think this is a great all-around solution, either. And it won't work when the resource is a service rather than a file. Move the service to the final web server/database I suppose I could run a web server on my SQL Server database, so the web site can connect to it using impersonation and make me happy. But do we really want random extra web servers on our database servers just so this is possible? No. Virtual directories in IIS I know that virtual directories can help make remote resources look as though they are local, and this supports using custom credentials for each virtual directory. I haven't been able to come up with, yet, how this would solve the problem for system calls. Users could reach file shares directly, but this won't help, say, classic ASP code access resources. I could use a URL instead of a file path to read remote data files in a web page, but this isn't going to help me make a connection to an Access database, a SQL server database, or any other resource that uses a connection library rather than being able to just read all the bytes and work with them. I wish there was some kind of "service tunnel" that I could create. Think about how a VPN makes remote resources look like they are local. With a richer aliasing mechanism, perhaps code-based, why couldn't even database connections occur under a defined security context? Why not a special Windows component that lets you specify, per user, what resources are available and what alternate credentials are used for the connection? File shares, databases, web sites, you name it. I guess I'm almost talking about a specialized local proxy server. Anyway, so there's my list. I may update it if I think of more. Does anyone have any ideas for me? My current problem today is, yet again, I need a web site to connect to an Access database on a file share. Here we go again...

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  • solved: puppet master REST API returns 403 when running under passenger works when master runs from command line

    - by Anadi Misra
    I am using the standard auth.conf provided in puppet install for the puppet master which is running through passenger under Nginx. However for most of the catalog, files and certitifcate request I get a 403 response. ### Authenticated paths - these apply only when the client ### has a valid certificate and is thus authenticated # allow nodes to retrieve their own catalog path ~ ^/catalog/([^/]+)$ method find allow $1 # allow nodes to retrieve their own node definition path ~ ^/node/([^/]+)$ method find allow $1 # allow all nodes to access the certificates services path ~ ^/certificate_revocation_list/ca method find allow * # allow all nodes to store their reports path /report method save allow * # unconditionally allow access to all file services # which means in practice that fileserver.conf will # still be used path /file allow * ### Unauthenticated ACL, for clients for which the current master doesn't ### have a valid certificate; we allow authenticated users, too, because ### there isn't a great harm in letting that request through. # allow access to the master CA path /certificate/ca auth any method find allow * path /certificate/ auth any method find allow * path /certificate_request auth any method find, save allow * path /facts auth any method find, search allow * # this one is not stricly necessary, but it has the merit # of showing the default policy, which is deny everything else path / auth any Puppet master however does not seems to be following this as I get this error on client [amisr1@blramisr195602 ~]$ sudo puppet agent --no-daemonize --verbose --server bangvmpllda02.XXXXX.com [sudo] password for amisr1: Starting Puppet client version 3.0.1 Warning: Unable to fetch my node definition, but the agent run will continue: Warning: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /certificate_revocation_list/ca [find] at :110 Info: Retrieving plugin Error: /File[/var/lib/puppet/lib]: Failed to generate additional resources using 'eval_generate: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /file_metadata/plugins [search] at :110 Error: /File[/var/lib/puppet/lib]: Could not evaluate: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /file_metadata/plugins [find] at :110 Could not retrieve file metadata for puppet://devops.XXXXX.com/plugins: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /file_metadata/plugins [find] at :110 Error: Could not retrieve catalog from remote server: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /catalog/blramisr195602.XXXXX.com [find] at :110 Using cached catalog Error: Could not retrieve catalog; skipping run Error: Could not send report: Error 403 on SERVER: Forbidden request: XX.XXX.XX.XX(XX.XXX.XX.XX) access to /report/blramisr195602.XXXXX.com [save] at :110 and the server logs show XX.XXX.XX.XX - - [10/Dec/2012:14:46:52 +0530] "GET /production/certificate_revocation_list/ca? HTTP/1.1" 403 102 "-" "Ruby" XX.XXX.XX.XX - - [10/Dec/2012:14:46:52 +0530] "GET /production/file_metadatas/plugins?links=manage&recurse=true&&ignore=---+%0A++-+%22.svn%22%0A++-+CVS%0A++-+%22.git%22&checksum_type=md5 HTTP/1.1" 403 95 "-" "Ruby" XX.XXX.XX.XX - - [10/Dec/2012:14:46:52 +0530] "GET /production/file_metadata/plugins? HTTP/1.1" 403 93 "-" "Ruby" XX.XXX.XX.XX - - [10/Dec/2012:14:46:53 +0530] "POST /production/catalog/blramisr195602.XXXXX.com HTTP/1.1" 403 106 "-" "Ruby" XX.XXX.XX.XX - - [10/Dec/2012:14:46:53 +0530] "PUT /production/report/blramisr195602.XXXXX.com HTTP/1.1" 403 105 "-" "Ruby" thefile server conf file is as follows (and goin by what they say on puppet site, It is better to regulate access in auth.conf for reaching file server and then allow file server to server all) [files] path /apps/puppet/files allow * [private] path /apps/puppet/private/%H allow * [modules] allow * I am using server and client version 3 Nginx has been compiled using the following options nginx version: nginx/1.3.9 built by gcc 4.4.6 20120305 (Red Hat 4.4.6-4) (GCC) TLS SNI support enabled configure arguments: --prefix=/apps/nginx --conf-path=/apps/nginx/nginx.conf --pid-path=/apps/nginx/run/nginx.pid --error-log-path=/apps/nginx/logs/error.log --http-log-path=/apps/nginx/logs/access.log --with-http_ssl_module --with-http_gzip_static_module --add-module=/usr/lib/ruby/gems/1.8/gems/passenger-3.0.18/ext/nginx --add-module=/apps/Downloads/nginx/nginx-auth-ldap-master/ and the standard nginx puppet master conf server { ssl on; listen 8140 ssl; server_name _; passenger_enabled on; passenger_set_cgi_param HTTP_X_CLIENT_DN $ssl_client_s_dn; passenger_set_cgi_param HTTP_X_CLIENT_VERIFY $ssl_client_verify; passenger_min_instances 5; access_log logs/puppet_access.log; error_log logs/puppet_error.log; root /apps/nginx/html/rack/public; ssl_certificate /var/lib/puppet/ssl/certs/bangvmpllda02.XXXXXX.com.pem; ssl_certificate_key /var/lib/puppet/ssl/private_keys/bangvmpllda02.XXXXXX.com.pem; ssl_crl /var/lib/puppet/ssl/ca/ca_crl.pem; ssl_client_certificate /var/lib/puppet/ssl/certs/ca.pem; ssl_ciphers SSLv2:-LOW:-EXPORT:RC4+RSA; ssl_prefer_server_ciphers on; ssl_verify_client optional; ssl_verify_depth 1; ssl_session_cache shared:SSL:128m; ssl_session_timeout 5m; } Puppet is picking up the correct settings from the files mentioned because config print command points to /etc/puppet [amisr1@bangvmpllDA02 puppet]$ sudo puppet config print | grep conf async_storeconfigs = false authconfig = /etc/puppet/namespaceauth.conf autosign = /etc/puppet/autosign.conf catalog_cache_terminus = store_configs confdir = /etc/puppet config = /etc/puppet/puppet.conf config_file_name = puppet.conf config_version = "" configprint = all configtimeout = 120 dblocation = /var/lib/puppet/state/clientconfigs.sqlite3 deviceconfig = /etc/puppet/device.conf fileserverconfig = /etc/puppet/fileserver.conf genconfig = false hiera_config = /etc/puppet/hiera.yaml localconfig = /var/lib/puppet/state/localconfig name = config rest_authconfig = /etc/puppet/auth.conf storeconfigs = true storeconfigs_backend = puppetdb tagmap = /etc/puppet/tagmail.conf thin_storeconfigs = false I checked the firewall rules on this VM; 80, 443, 8140, 3000 are allowed. Do I still have to tweak any specifics to auth.conf for getting this to work? Update I added verbose logging to the puppet master and restarted nginx; here's the additional info I see in logs Mon Dec 10 18:19:15 +0530 2012 Puppet (err): Could not resolve 10.209.47.31: no name for 10.209.47.31 Mon Dec 10 18:19:15 +0530 2012 access[/] (info): defaulting to no access for 10.209.47.31 Mon Dec 10 18:19:15 +0530 2012 Puppet (warning): Denying access: Forbidden request: 10.209.47.31(10.209.47.31) access to /file_metadata/plugins [find] at :111 Mon Dec 10 18:19:15 +0530 2012 Puppet (err): Forbidden request: 10.209.47.31(10.209.47.31) access to /file_metadata/plugins [find] at :111 10.209.47.31 - - [10/Dec/2012:18:19:15 +0530] "GET /production/file_metadata/plugins? HTTP/1.1" 403 93 "-" "Ruby" On the agent machine facter fqdn and hostname both return a fully qualified host name [amisr1@blramisr195602 ~]$ sudo facter fqdn blramisr195602.XXXXXXX.com I then updated the agent configuration to add dns_alt_names = 10.209.47.31 cleaned all certificates on master and agent and regenerated the certificates and signed them on master using the option --allow-dns-alt-names [amisr1@bangvmpllDA02 ~]$ sudo puppet cert sign blramisr195602.XXXXXX.com Error: CSR 'blramisr195602.XXXXXX.com' contains subject alternative names (DNS:10.209.47.31, DNS:blramisr195602.XXXXXX.com), which are disallowed. Use `puppet cert --allow-dns-alt-names sign blramisr195602.XXXXXX.com` to sign this request. [amisr1@bangvmpllDA02 ~]$ sudo puppet cert --allow-dns-alt-names sign blramisr195602.XXXXXX.com Signed certificate request for blramisr195602.XXXXXX.com Removing file Puppet::SSL::CertificateRequest blramisr195602.XXXXXX.com at '/var/lib/puppet/ssl/ca/requests/blramisr195602.XXXXXX.com.pem' however, that doesn't help either; I get same errors as before. Not sure why in the logs it shows comparing access rules by IP and not hostname. Is there any Nginx configuration to change this behavior?

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  • ColdFusion Server crash after thousands of HTTP requests

    - by Jason Bristol
    We are running ColdFusion 8 on a windows server 2003 VPS with an API that exposes student records to a partner API through a connector. Our API returns around 50k student records serialized in XML format pretty seamlessly. My question originates when something very frightening happened today when we tested our connector to our partners API. Our entire website and web host went down. We assumed that our host was just having some issues and after 4 hours with no resolution and no response from their customer service we finally got a response from them claiming that they had an "unauthorized user" in their network. After our server was back up we were unable to connect to our website as if the web service or coldfusion itself had froze. This is really where my concern comes from as I fear we may have overloaded the web service. As I mentioned before we tried sending over 50k HTTP POST requests over to our partner's API, however everything stopped after around 1.6k Is this bad practice or is there some sort of rate limiting I can relax somewhere in server configuration? We managed to find a workaround, but it bypasses our connector which is critical to our design. This would have been a one time deal as the purpose of so many requests was to populate our partner's website with current data, after that hourly syncs will keep requests down to around 100 per hour. UPDATE Our partner API is owned and operated by Pardot. We are converting students to prospects by passing student data to their API which unfortunately only seems to accept one student at a time. For that reason we have to do all 50k requests individually. Our server has 4GB of RAM, an Intel Core 2 Duo @ 2.8GHz running Windows Server 2003 SP2. I monitored the server during a 100 student sync, a 400 student sync, and a 1.4k student sync with the following results: 100 students - 2.25GB of Memory, 30-40% CPU utilization, 0.2-0.3% network bandwidth 400 students - 2.30GB of Memory, 30-50% CPU utilization, 0.2-1.0% network bandwidth 1.4k students - 2.30GB of Memory, 30-70% CPU utilization, 0.2-1.0% network bandwidth I know this is a far cry from 50k students, but I don't want to risk taking down our CMS system again assuming that was the cause. To give you a look at our code: <cfif (#getStudents.statusCode# eq "200 OK")> <cftry> <cfloop index="StudentXML" array="#XmlSearch(responseSTUD,'/students/student')#"> <cfset StudentXML = XmlParse(StudentXML)> <cfhttp url="#PARDOT_CMS_UPSERT#" method="post" timeout="10000" > <cfhttpparam type="url" name="user_key" value="#PARDOT_CMS_USERKEY#"> <cfhttpparam type="url" name="api_key" value="#api_key#"> <cfhttpparam type="url" name="email" value="#StudentXML.student.email.XmlText#"> <cfhttpparam type="url" name="first_name" value="#StudentXML.student.first.XmlText#"> <cfhttpparam type="url" name="last_name" value="#StudentXML.student.last.XmlText#"> <cfhttpparam type="url" name="in_cms" value="#StudentXML.student.studentid.XmlText#"> <cfhttpparam type="url" name="company" value="#StudentXML.student.agencyname.XmlText#"> <cfhttpparam type="url" name="country" value="#StudentXML.student.countryname.XmlText#"> <cfhttpparam type="url" name="address_one" value="#StudentXML.student.address.XmlText#"> <cfhttpparam type="url" name="address_two" value="#StudentXML.student.address2.XmlText#"> <cfhttpparam type="url" name="city" value="#StudentXML.student.city.XmlText#"> <cfhttpparam type="url" name="state" value="#StudentXML.student.state_province.XmlText#"> <cfhttpparam type="url" name="zip" value="#StudentXML.student.postalcode.XmlText#"> <cfhttpparam type="url" name="phone" value="#StudentXML.student.phone.XmlText#"> <cfhttpparam type="url" name="fax" value="#StudentXML.student.fax.XmlText#"> <cfhttpparam type="url" name="output" value="simple"> </cfhttp> </cfloop> <cfcatch type="any"> <cfdump var="#cfcatch.Message#"> </cfcatch> </cftry> </cfif> UPDATE 2 I checked the CF logs and found a couple of these: "Error","jrpp-8","06/06/13","16:10:18","CMS-API","Java heap space The specific sequence of files included or processed is: D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm, line: 675 " java.lang.OutOfMemoryError: Java heap space at java.util.Arrays.copyOf(Arrays.java:2882) at java.io.CharArrayWriter.write(CharArrayWriter.java:105) at coldfusion.runtime.CharBuffer.replace(CharBuffer.java:37) at coldfusion.runtime.CharBuffer.replace(CharBuffer.java:50) at coldfusion.runtime.NeoBodyContent.write(NeoBodyContent.java:254) at cfapi2ecfm292155732._factor30(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:675) at cfapi2ecfm292155732._factor31(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:662) at cfapi2ecfm292155732._factor36(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:659) at cfapi2ecfm292155732._factor42(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:657) at cfapi2ecfm292155732._factor37(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm) at cfapi2ecfm292155732._factor44(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:456) at cfapi2ecfm292155732._factor38(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm) at cfapi2ecfm292155732._factor46(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:455) at cfapi2ecfm292155732._factor39(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm) at cfapi2ecfm292155732._factor47(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:453) at cfapi2ecfm292155732.runPage(D:\Clients\www.xxx.com\www\dev.cms\api\v1\api.cfm:1) at coldfusion.runtime.CfJspPage.invoke(CfJspPage.java:192) at coldfusion.tagext.lang.IncludeTag.doStartTag(IncludeTag.java:366) at coldfusion.filter.CfincludeFilter.invoke(CfincludeFilter.java:65) at coldfusion.filter.ApplicationFilter.invoke(ApplicationFilter.java:279) at coldfusion.filter.RequestMonitorFilter.invoke(RequestMonitorFilter.java:48) at coldfusion.filter.MonitoringFilter.invoke(MonitoringFilter.java:40) at coldfusion.filter.PathFilter.invoke(PathFilter.java:86) at coldfusion.filter.ExceptionFilter.invoke(ExceptionFilter.java:70) at coldfusion.filter.ClientScopePersistenceFilter.invoke(ClientScopePersistenceFilter.java:28) at coldfusion.filter.BrowserFilter.invoke(BrowserFilter.java:38) at coldfusion.filter.NoCacheFilter.invoke(NoCacheFilter.java:46) at coldfusion.filter.GlobalsFilter.invoke(GlobalsFilter.java:38) at coldfusion.filter.DatasourceFilter.invoke(DatasourceFilter.java:22) at coldfusion.CfmServlet.service(CfmServlet.java:175) at coldfusion.bootstrap.BootstrapServlet.service(BootstrapServlet.java:89) at jrun.servlet.FilterChain.doFilter(FilterChain.java:86) Looks like I might have crashed the JVM in CF, is there a better way to do this? We are thinking of just exporting all records initially as a CSV file and importing it into Pardot seeing as we will never have to do a request this large again.

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  • IIS Strategies for Accessing Secured Network Resources

    - by Emtucifor
    Problem: A user connects to a service on a machine, such as an IIS web site or a SQL Server database. The site or the database need to gain access to network resources such as file shares (the most common) or a database on a different server. Permission is denied. This is because the user the service is running as doesn't have network permissions in the first place, or if it does, it doesn't have rights to access the remote resource. I keep running into this problem over and over again and am tired of not having a really solid way of handling it. Here are some workarounds I'm aware of: Run IIS as a custom-created domain user who is granted high permissions If permissions are granted one file share at a time, then every time I want to read from a new share, I would have to ask a network admin to add it for me. Eventually, with many web sites reading from many shares, it is going to get really complicated. If permissions are just opened up wide for the user to access any file shares in our domain, then this seems like an unnecessary security surface area to present. This also applies to all the sites running on IIS, rather than just the selected site or virtual directory that needs the access, a further surface area problem. Still use the IUSR account but give it network permissions and set up the same user name on the remote resource (not a domain user, a local user) This also has its problems. For example, there's a file share I am using that I have full rights to for sharing, but I can't log in to the machine. So I have to find the right admin and ask him to do it for me. Any time something has to change, it's another request to an admin. Allow IIS users to connect as anonymous, but set the account used for anonymous access to a high-privilege one This is even worse than giving the IIS IUSR full privileges, because it means my web site can't use any kind of security in the first place. Connect using Kerberos, then delegate This sounds good in principle but has all sorts of problems. First of all, if you're using virtual web sites where the domain name you connect to the site with is not the base machine name (as we do frequently), then you have to set up a Service Principal Name on the webserver using Microsoft's SetSPN utility. It's complicated and apparently prone to errors. Also, you have to ask your network/domain admin to change security policy for the web server so it is "trusted for delegation." If you don't get everything perfectly right, suddenly your intended Kerberos authentication is NTLM instead, and you can only impersonate rather than delegate, and thus no reaching out over the network as the user. Also, this method can be problematic because sometimes you need the web site or database to have permissions that the connecting user doesn't have. Create a service or COM+ application that fetches the resource for the web site Services and COM+ packages are run with their own set of credentials. Running as a high-privilege user is okay since they can do their own security and deny requests that are not legitimate, putting control in the hands of the application developer instead of the network admin. Problems: I am using a COM+ package that does exactly this on Windows Server 2000 to deliver highly sensitive images to a secured web application. I tried moving the web site to Windows Server 2003 and was suddenly denied permission to instantiate the COM+ object, very likely registry permissions. I trolled around quite a bit and did not solve the problem, partly because I was reluctant to give the IUSR account full registry permissions. That seems like the same bad practice as just running IIS as a high-privilege user. Note: This is actually really simple. In a programming language of your choice, you create a class with a function that returns an instance of the object you want (an ADODB.Connection, for example), and build a dll, which you register as a COM+ object. In your web server-side code, you create an instance of the class and use the function, and since it is running under a different security context, calls to network resources work. Map drive letters to shares This could theoretically work, but in my mind it's not really a good long-term strategy. Even though mappings can be created with specific credentials, and this can be done by others than a network admin, this also is going to mean that there are either way too many shared drives (small granularity) or too much permission is granted to entire file servers (large granularity). Also, I haven't figured out how to map a drive so that the IUSR gets the drives. Mapping a drive is for the current user, I don't know the IUSR account password to log in as it and create the mappings. Move the resources local to the web server/database There are times when I've done this, especially with Access databases. Does the database have to live out on the file share? Sometimes, it was just easiest to move the database to the web server or to the SQL database server (so the linked server to it would work). But I don't think this is a great all-around solution, either. And it won't work when the resource is a service rather than a file. Move the service to the final web server/database I suppose I could run a web server on my SQL Server database, so the web site can connect to it using impersonation and make me happy. But do we really want random extra web servers on our database servers just so this is possible? No. Virtual directories in IIS I know that virtual directories can help make remote resources look as though they are local, and this supports using custom credentials for each virtual directory. I haven't been able to come up with, yet, how this would solve the problem for system calls. Users could reach file shares directly, but this won't help, say, classic ASP code access resources. I could use a URL instead of a file path to read remote data files in a web page, but this isn't going to help me make a connection to an Access database, a SQL server database, or any other resource that uses a connection library rather than being able to just read all the bytes and work with them. I wish there was some kind of "service tunnel" that I could create. Think about how a VPN makes remote resources look like they are local. With a richer aliasing mechanism, perhaps code-based, why couldn't even database connections occur under a defined security context? Why not a special Windows component that lets you specify, per user, what resources are available and what alternate credentials are used for the connection? File shares, databases, web sites, you name it. I guess I'm almost talking about a specialized local proxy server. Anyway, so there's my list. I may update it if I think of more. Does anyone have any ideas for me? My current problem today is, yet again, I need a web site to connect to an Access database on a file share. Here we go again...

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  • SimpleMembership, Membership Providers, Universal Providers and the new ASP.NET 4.5 Web Forms and ASP.NET MVC 4 templates

    - by Jon Galloway
    The ASP.NET MVC 4 Internet template adds some new, very useful features which are built on top of SimpleMembership. These changes add some great features, like a much simpler and extensible membership API and support for OAuth. However, the new account management features require SimpleMembership and won't work against existing ASP.NET Membership Providers. I'll start with a summary of top things you need to know, then dig into a lot more detail. Summary: SimpleMembership has been designed as a replacement for traditional the previous ASP.NET Role and Membership provider system SimpleMembership solves common problems people ran into with the Membership provider system and was designed for modern user / membership / storage needs SimpleMembership integrates with the previous membership system, but you can't use a MembershipProvider with SimpleMembership The new ASP.NET MVC 4 Internet application template AccountController requires SimpleMembership and is not compatible with previous MembershipProviders You can continue to use existing ASP.NET Role and Membership providers in ASP.NET 4.5 and ASP.NET MVC 4 - just not with the ASP.NET MVC 4 AccountController The existing ASP.NET Role and Membership provider system remains supported as is part of the ASP.NET core ASP.NET 4.5 Web Forms does not use SimpleMembership; it implements OAuth on top of ASP.NET Membership The ASP.NET Web Site Administration Tool (WSAT) is not compatible with SimpleMembership The following is the result of a few conversations with Erik Porter (PM for ASP.NET MVC) to make sure I had some the overall details straight, combined with a lot of time digging around in ILSpy and Visual Studio's assembly browsing tools. SimpleMembership: The future of membership for ASP.NET The ASP.NET Membership system was introduces with ASP.NET 2.0 back in 2005. It was designed to solve common site membership requirements at the time, which generally involved username / password based registration and profile storage in SQL Server. It was designed with a few extensibility mechanisms - notably a provider system (which allowed you override some specifics like backing storage) and the ability to store additional profile information (although the additional  profile information was packed into a single column which usually required access through the API). While it's sometimes frustrating to work with, it's held up for seven years - probably since it handles the main use case (username / password based membership in a SQL Server database) smoothly and can be adapted to most other needs (again, often frustrating, but it can work). The ASP.NET Web Pages and WebMatrix efforts allowed the team an opportunity to take a new look at a lot of things - e.g. the Razor syntax started with ASP.NET Web Pages, not ASP.NET MVC. The ASP.NET Web Pages team designed SimpleMembership to (wait for it) simplify the task of dealing with membership. As Matthew Osborn said in his post Using SimpleMembership With ASP.NET WebPages: With the introduction of ASP.NET WebPages and the WebMatrix stack our team has really be focusing on making things simpler for the developer. Based on a lot of customer feedback one of the areas that we wanted to improve was the built in security in ASP.NET. So with this release we took that time to create a new built in (and default for ASP.NET WebPages) security provider. I say provider because the new stuff is still built on the existing ASP.NET framework. So what do we call this new hotness that we have created? Well, none other than SimpleMembership. SimpleMembership is an umbrella term for both SimpleMembership and SimpleRoles. Part of simplifying membership involved fixing some common problems with ASP.NET Membership. Problems with ASP.NET Membership ASP.NET Membership was very obviously designed around a set of assumptions: Users and user information would most likely be stored in a full SQL Server database or in Active Directory User and profile information would be optimized around a set of common attributes (UserName, Password, IsApproved, CreationDate, Comment, Role membership...) and other user profile information would be accessed through a profile provider Some problems fall out of these assumptions. Requires Full SQL Server for default cases The default, and most fully featured providers ASP.NET Membership providers (SQL Membership Provider, SQL Role Provider, SQL Profile Provider) require full SQL Server. They depend on stored procedure support, and they rely on SQL Server cache dependencies, they depend on agents for clean up and maintenance. So the main SQL Server based providers don't work well on SQL Server CE, won't work out of the box on SQL Azure, etc. Note: Cory Fowler recently let me know about these Updated ASP.net scripts for use with Microsoft SQL Azure which do support membership, personalization, profile, and roles. But the fact that we need a support page with a set of separate SQL scripts underscores the underlying problem. Aha, you say! Jon's forgetting the Universal Providers, a.k.a. System.Web.Providers! Hold on a bit, we'll get to those... Custom Membership Providers have to work with a SQL-Server-centric API If you want to work with another database or other membership storage system, you need to to inherit from the provider base classes and override a bunch of methods which are tightly focused on storing a MembershipUser in a relational database. It can be done (and you can often find pretty good ones that have already been written), but it's a good amount of work and often leaves you with ugly code that has a bunch of System.NotImplementedException fun since there are a lot of methods that just don't apply. Designed around a specific view of users, roles and profiles The existing providers are focused on traditional membership - a user has a username and a password, some specific roles on the site (e.g. administrator, premium user), and may have some additional "nice to have" optional information that can be accessed via an API in your application. This doesn't fit well with some modern usage patterns: In OAuth and OpenID, the user doesn't have a password Often these kinds of scenarios map better to user claims or rights instead of monolithic user roles For many sites, profile or other non-traditional information is very important and needs to come from somewhere other than an API call that maps to a database blob What would work a lot better here is a system in which you were able to define your users, rights, and other attributes however you wanted and the membership system worked with your model - not the other way around. Requires specific schema, overflow in blob columns I've already mentioned this a few times, but it bears calling out separately - ASP.NET Membership focuses on SQL Server storage, and that storage is based on a very specific database schema. SimpleMembership as a better membership system As you might have guessed, SimpleMembership was designed to address the above problems. Works with your Schema As Matthew Osborn explains in his Using SimpleMembership With ASP.NET WebPages post, SimpleMembership is designed to integrate with your database schema: All SimpleMembership requires is that there are two columns on your users table so that we can hook up to it – an “ID” column and a “username” column. The important part here is that they can be named whatever you want. For instance username doesn't have to be an alias it could be an email column you just have to tell SimpleMembership to treat that as the “username” used to log in. Matthew's example shows using a very simple user table named Users (it could be named anything) with a UserID and Username column, then a bunch of other columns he wanted in his app. Then we point SimpleMemberhip at that table with a one-liner: WebSecurity.InitializeDatabaseFile("SecurityDemo.sdf", "Users", "UserID", "Username", true); No other tables are needed, the table can be named anything we want, and can have pretty much any schema we want as long as we've got an ID and something that we can map to a username. Broaden database support to the whole SQL Server family While SimpleMembership is not database agnostic, it works across the SQL Server family. It continues to support full SQL Server, but it also works with SQL Azure, SQL Server CE, SQL Server Express, and LocalDB. Everything's implemented as SQL calls rather than requiring stored procedures, views, agents, and change notifications. Note that SimpleMembership still requires some flavor of SQL Server - it won't work with MySQL, NoSQL databases, etc. You can take a look at the code in WebMatrix.WebData.dll using a tool like ILSpy if you'd like to see why - there places where SQL Server specific SQL statements are being executed, especially when creating and initializing tables. It seems like you might be able to work with another database if you created the tables separately, but I haven't tried it and it's not supported at this point. Note: I'm thinking it would be possible for SimpleMembership (or something compatible) to run Entity Framework so it would work with any database EF supports. That seems useful to me - thoughts? Note: SimpleMembership has the same database support - anything in the SQL Server family - that Universal Providers brings to the ASP.NET Membership system. Easy to with Entity Framework Code First The problem with with ASP.NET Membership's system for storing additional account information is that it's the gate keeper. That means you're stuck with its schema and accessing profile information through its API. SimpleMembership flips that around by allowing you to use any table as a user store. That means you're in control of the user profile information, and you can access it however you'd like - it's just data. Let's look at a practical based on the AccountModel.cs class in an ASP.NET MVC 4 Internet project. Here I'm adding a Birthday property to the UserProfile class. [Table("UserProfile")] public class UserProfile { [Key] [DatabaseGeneratedAttribute(DatabaseGeneratedOption.Identity)] public int UserId { get; set; } public string UserName { get; set; } public DateTime Birthday { get; set; } } Now if I want to access that information, I can just grab the account by username and read the value. var context = new UsersContext(); var username = User.Identity.Name; var user = context.UserProfiles.SingleOrDefault(u => u.UserName == username); var birthday = user.Birthday; So instead of thinking of SimpleMembership as a big membership API, think of it as something that handles membership based on your user database. In SimpleMembership, everything's keyed off a user row in a table you define rather than a bunch of entries in membership tables that were out of your control. How SimpleMembership integrates with ASP.NET Membership Okay, enough sales pitch (and hopefully background) on why things have changed. How does this affect you? Let's start with a diagram to show the relationship (note: I've simplified by removing a few classes to show the important relationships): So SimpleMembershipProvider is an implementaiton of an ExtendedMembershipProvider, which inherits from MembershipProvider and adds some other account / OAuth related things. Here's what ExtendedMembershipProvider adds to MembershipProvider: The important thing to take away here is that a SimpleMembershipProvider is a MembershipProvider, but a MembershipProvider is not a SimpleMembershipProvider. This distinction is important in practice: you cannot use an existing MembershipProvider (including the Universal Providers found in System.Web.Providers) with an API that requires a SimpleMembershipProvider, including any of the calls in WebMatrix.WebData.WebSecurity or Microsoft.Web.WebPages.OAuth.OAuthWebSecurity. However, that's as far as it goes. Membership Providers still work if you're accessing them through the standard Membership API, and all of the core stuff  - including the AuthorizeAttribute, role enforcement, etc. - will work just fine and without any change. Let's look at how that affects you in terms of the new templates. Membership in the ASP.NET MVC 4 project templates ASP.NET MVC 4 offers six Project Templates: Empty - Really empty, just the assemblies, folder structure and a tiny bit of basic configuration. Basic - Like Empty, but with a bit of UI preconfigured (css / images / bundling). Internet - This has both a Home and Account controller and associated views. The Account Controller supports registration and login via either local accounts and via OAuth / OpenID providers. Intranet - Like the Internet template, but it's preconfigured for Windows Authentication. Mobile - This is preconfigured using jQuery Mobile and is intended for mobile-only sites. Web API - This is preconfigured for a service backend built on ASP.NET Web API. Out of these templates, only one (the Internet template) uses SimpleMembership. ASP.NET MVC 4 Basic template The Basic template has configuration in place to use ASP.NET Membership with the Universal Providers. You can see that configuration in the ASP.NET MVC 4 Basic template's web.config: <profile defaultProvider="DefaultProfileProvider"> <providers> <add name="DefaultProfileProvider" type="System.Web.Providers.DefaultProfileProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </profile> <membership defaultProvider="DefaultMembershipProvider"> <providers> <add name="DefaultMembershipProvider" type="System.Web.Providers.DefaultMembershipProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" enablePasswordRetrieval="false" enablePasswordReset="true" requiresQuestionAndAnswer="false" requiresUniqueEmail="false" maxInvalidPasswordAttempts="5" minRequiredPasswordLength="6" minRequiredNonalphanumericCharacters="0" passwordAttemptWindow="10" applicationName="/" /> </providers> </membership> <roleManager defaultProvider="DefaultRoleProvider"> <providers> <add name="DefaultRoleProvider" type="System.Web.Providers.DefaultRoleProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </roleManager> <sessionState mode="InProc" customProvider="DefaultSessionProvider"> <providers> <add name="DefaultSessionProvider" type="System.Web.Providers.DefaultSessionStateProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" /> </providers> </sessionState> This means that it's business as usual for the Basic template as far as ASP.NET Membership works. ASP.NET MVC 4 Internet template The Internet template has a few things set up to bootstrap SimpleMembership: \Models\AccountModels.cs defines a basic user account and includes data annotations to define keys and such \Filters\InitializeSimpleMembershipAttribute.cs creates the membership database using the above model, then calls WebSecurity.InitializeDatabaseConnection which verifies that the underlying tables are in place and marks initialization as complete (for the application's lifetime) \Controllers\AccountController.cs makes heavy use of OAuthWebSecurity (for OAuth account registration / login / management) and WebSecurity. WebSecurity provides account management services for ASP.NET MVC (and Web Pages) WebSecurity can work with any ExtendedMembershipProvider. There's one in the box (SimpleMembershipProvider) but you can write your own. Since a standard MembershipProvider is not an ExtendedMembershipProvider, WebSecurity will throw exceptions if the default membership provider is a MembershipProvider rather than an ExtendedMembershipProvider. Practical example: Create a new ASP.NET MVC 4 application using the Internet application template Install the Microsoft ASP.NET Universal Providers for LocalDB NuGet package Run the application, click on Register, add a username and password, and click submit You'll get the following execption in AccountController.cs::Register: To call this method, the "Membership.Provider" property must be an instance of "ExtendedMembershipProvider". This occurs because the ASP.NET Universal Providers packages include a web.config transform that will update your web.config to add the Universal Provider configuration I showed in the Basic template example above. When WebSecurity tries to use the configured ASP.NET Membership Provider, it checks if it can be cast to an ExtendedMembershipProvider before doing anything else. So, what do you do? Options: If you want to use the new AccountController, you'll either need to use the SimpleMembershipProvider or another valid ExtendedMembershipProvider. This is pretty straightforward. If you want to use an existing ASP.NET Membership Provider in ASP.NET MVC 4, you can't use the new AccountController. You can do a few things: Replace  the AccountController.cs and AccountModels.cs in an ASP.NET MVC 4 Internet project with one from an ASP.NET MVC 3 application (you of course won't have OAuth support). Then, if you want, you can go through and remove other things that were built around SimpleMembership - the OAuth partial view, the NuGet packages (e.g. the DotNetOpenAuthAuth package, etc.) Use an ASP.NET MVC 4 Internet application template and add in a Universal Providers NuGet package. Then copy in the AccountController and AccountModel classes. Create an ASP.NET MVC 3 project and upgrade it to ASP.NET MVC 4 using the steps shown in the ASP.NET MVC 4 release notes. None of these are particularly elegant or simple. Maybe we (or just me?) can do something to make this simpler - perhaps a NuGet package. However, this should be an edge case - hopefully the cases where you'd need to create a new ASP.NET but use legacy ASP.NET Membership Providers should be pretty rare. Please let me (or, preferably the team) know if that's an incorrect assumption. Membership in the ASP.NET 4.5 project template ASP.NET 4.5 Web Forms took a different approach which builds off ASP.NET Membership. Instead of using the WebMatrix security assemblies, Web Forms uses Microsoft.AspNet.Membership.OpenAuth assembly. I'm no expert on this, but from a bit of time in ILSpy and Visual Studio's (very pretty) dependency graphs, this uses a Membership Adapter to save OAuth data into an EF managed database while still running on top of ASP.NET Membership. Note: There may be a way to use this in ASP.NET MVC 4, although it would probably take some plumbing work to hook it up. How does this fit in with Universal Providers (System.Web.Providers)? Just to summarize: Universal Providers are intended for cases where you have an existing ASP.NET Membership Provider and you want to use it with another SQL Server database backend (other than SQL Server). It doesn't require agents to handle expired session cleanup and other background tasks, it piggybacks these tasks on other calls. Universal Providers are not really, strictly speaking, universal - at least to my way of thinking. They only work with databases in the SQL Server family. Universal Providers do not work with Simple Membership. The Universal Providers packages include some web config transforms which you would normally want when you're using them. What about the Web Site Administration Tool? Visual Studio includes tooling to launch the Web Site Administration Tool (WSAT) to configure users and roles in your application. WSAT is built to work with ASP.NET Membership, and is not compatible with Simple Membership. There are two main options there: Use the WebSecurity and OAuthWebSecurity API to manage the users and roles Create a web admin using the above APIs Since SimpleMembership runs on top of your database, you can update your users as you would any other data - via EF or even in direct database edits (in development, of course)

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  • Rendering ASP.NET Script References into the Html Header

    - by Rick Strahl
    One thing that I’ve come to appreciate in control development in ASP.NET that use JavaScript is the ability to have more control over script and script include placement than ASP.NET provides natively. Specifically in ASP.NET you can use either the ClientScriptManager or ScriptManager to embed scripts and script references into pages via code. This works reasonably well, but the script references that get generated are generated into the HTML body and there’s very little operational control for placement of scripts. If you have multiple controls or several of the same control that need to place the same scripts onto the page it’s not difficult to end up with scripts that render in the wrong order and stop working correctly. This is especially critical if you load script libraries with dependencies either via resources or even if you are rendering referenced to CDN resources. Natively ASP.NET provides a host of methods that help embedding scripts into the page via either Page.ClientScript or the ASP.NET ScriptManager control (both with slightly different syntax): RegisterClientScriptBlock Renders a script block at the top of the HTML body and should be used for embedding callable functions/classes. RegisterStartupScript Renders a script block just prior to the </form> tag and should be used to for embedding code that should execute when the page is first loaded. Not recommended – use jQuery.ready() or equivalent load time routines. RegisterClientScriptInclude Embeds a reference to a script from a url into the page. RegisterClientScriptResource Embeds a reference to a Script from a resource file generating a long resource file string All 4 of these methods render their <script> tags into the HTML body. The script blocks give you a little bit of control by having a ‘top’ and ‘bottom’ of the document location which gives you some flexibility over script placement and precedence. Script includes and resource url unfortunately do not even get that much control – references are simply rendered into the page in the order of declaration. The ASP.NET ScriptManager control facilitates this task a little bit with the abililty to specify scripts in code and the ability to programmatically check what scripts have already been registered, but it doesn’t provide any more control over the script rendering process itself. Further the ScriptManager is a bear to deal with generically because generic code has to always check and see if it is actually present. Some time ago I posted a ClientScriptProxy class that helps with managing the latter process of sending script references either to ClientScript or ScriptManager if it’s available. Since I last posted about this there have been a number of improvements in this API, one of which is the ability to control placement of scripts and script includes in the page which I think is rather important and a missing feature in the ASP.NET native functionality. Handling ScriptRenderModes One of the big enhancements that I’ve come to rely on is the ability of the various script rendering functions described above to support rendering in multiple locations: /// <summary> /// Determines how scripts are included into the page /// </summary> public enum ScriptRenderModes { /// <summary> /// Inherits the setting from the control or from the ClientScript.DefaultScriptRenderMode /// </summary> Inherit, /// Renders the script include at the location of the control /// </summary> Inline, /// <summary> /// Renders the script include into the bottom of the header of the page /// </summary> Header, /// <summary> /// Renders the script include into the top of the header of the page /// </summary> HeaderTop, /// <summary> /// Uses ClientScript or ScriptManager to embed the script include to /// provide standard ASP.NET style rendering in the HTML body. /// </summary> Script, /// <summary> /// Renders script at the bottom of the page before the last Page.Controls /// literal control. Note this may result in unexpected behavior /// if /body and /html are not the last thing in the markup page. /// </summary> BottomOfPage } This enum is then applied to the various Register functions to allow more control over where scripts actually show up. Why is this useful? For me I often render scripts out of control resources and these scripts often include things like a JavaScript Library (jquery) and a few plug-ins. The order in which these can be loaded is critical so that jQuery.js always loads before any plug-in for example. Typically I end up with a general script layout like this: Core Libraries- HeaderTop Plug-ins: Header ScriptBlocks: Header or Script depending on other dependencies There’s also an option to render scripts and CSS at the very bottom of the page before the last Page control on the page which can be useful for speeding up page load when lots of scripts are loaded. The API syntax of the ClientScriptProxy methods is closely compatible with ScriptManager’s using static methods and control references to gain access to the page and embedding scripts. For example, to render some script into the current page in the header: // Create script block in header ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "hello_function", "function helloWorld() { alert('hello'); }", true, ScriptRenderModes.Header); // Same again - shouldn't be rendered because it's the same id ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "hello_function", "function helloWorld() { alert('hello'); }", true, ScriptRenderModes.Header); // Create a second script block in header ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "hello_function2", "function helloWorld2() { alert('hello2'); }", true, ScriptRenderModes.Header); // This just calls ClientScript and renders into bottom of document ClientScriptProxy.Current.RegisterStartupScript(this,typeof(ControlResources), "call_hello", "helloWorld();helloWorld2();", true); which generates: <html xmlns="http://www.w3.org/1999/xhtml" > <head><title> </title> <script type="text/javascript"> function helloWorld() { alert('hello'); } </script> <script type="text/javascript"> function helloWorld2() { alert('hello2'); } </script> </head> <body> … <script type="text/javascript"> //<![CDATA[ helloWorld();helloWorld2();//]]> </script> </form> </body> </html> Note that the scripts are generated into the header rather than the body except for the last script block which is the call to RegisterStartupScript. In general I wouldn’t recommend using RegisterStartupScript – ever. It’s a much better practice to use a script base load event to handle ‘startup’ code that should fire when the page first loads. So instead of the code above I’d actually recommend doing: ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "call_hello", "$().ready( function() { alert('hello2'); });", true, ScriptRenderModes.Header); assuming you’re using jQuery on the page. For script includes from a Url the following demonstrates how to embed scripts into the header. This example injects a jQuery and jQuery.UI script reference from the Google CDN then checks each with a script block to ensure that it has loaded and if not loads it from a server local location: // load jquery from CDN ClientScriptProxy.Current.RegisterClientScriptInclude(this, typeof(ControlResources), "http://ajax.googleapis.com/ajax/libs/jquery/1.3.2/jquery.min.js", ScriptRenderModes.HeaderTop); // check if jquery loaded - if it didn't we're not online string scriptCheck = @"if (typeof jQuery != 'object') document.write(unescape(""%3Cscript src='{0}' type='text/javascript'%3E%3C/script%3E""));"; string jQueryUrl = ClientScriptProxy.Current.GetWebResourceUrl(this, typeof(ControlResources), ControlResources.JQUERY_SCRIPT_RESOURCE); ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "jquery_register", string.Format(scriptCheck,jQueryUrl),true, ScriptRenderModes.HeaderTop); // Load jquery-ui from cdn ClientScriptProxy.Current.RegisterClientScriptInclude(this, typeof(ControlResources), "http://ajax.googleapis.com/ajax/libs/jqueryui/1.7.2/jquery-ui.min.js", ScriptRenderModes.Header); // check if we need to load from local string jQueryUiUrl = ResolveUrl("~/scripts/jquery-ui-custom.min.js"); ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "jqueryui_register", string.Format(scriptCheck, jQueryUiUrl), true, ScriptRenderModes.Header); // Create script block in header ClientScriptProxy.Current.RegisterClientScriptBlock(this, typeof(ControlResources), "hello_function", "$().ready( function() { alert('hello'); });", true, ScriptRenderModes.Header); which in turn generates this HTML: <html xmlns="http://www.w3.org/1999/xhtml" > <head> <script src="http://ajax.googleapis.com/ajax/libs/jquery/1.3.2/jquery.min.js" type="text/javascript"></script> <script type="text/javascript"> if (typeof jQuery != 'object') document.write(unescape("%3Cscript src='/WestWindWebToolkitWeb/WebResource.axd?d=DIykvYhJ_oXCr-TA_dr35i4AayJoV1mgnQAQGPaZsoPM2LCdvoD3cIsRRitHKlKJfV5K_jQvylK7tsqO3lQIFw2&t=633979863959332352' type='text/javascript'%3E%3C/script%3E")); </script> <title> </title> <script src="http://ajax.googleapis.com/ajax/libs/jqueryui/1.7.2/jquery-ui.min.js" type="text/javascript"></script> <script type="text/javascript"> if (typeof jQuery != 'object') document.write(unescape("%3Cscript src='/WestWindWebToolkitWeb/scripts/jquery-ui-custom.min.js' type='text/javascript'%3E%3C/script%3E")); </script> <script type="text/javascript"> $().ready(function() { alert('hello'); }); </script> </head> <body> …</body> </html> As you can see there’s a bit more control in this process as you can inject both script includes and script blocks into the document at the top or bottom of the header, plus if necessary at the usual body locations. This is quite useful especially if you create custom server controls that interoperate with script and have certain dependencies. The above is a good example of a useful switchable routine where you can switch where scripts load from by default – the above pulls from Google CDN but a configuration switch may automatically switch to pull from the local development copies if your doing development for example. How does it work? As mentioned the ClientScriptProxy object mimicks many of the ScriptManager script related methods and so provides close API compatibility with it although it contains many additional overloads that enhance functionality. It does however work against ScriptManager if it’s available on the page, or Page.ClientScript if it’s not so it provides a single unified frontend to script access. There are however many overloads of the original SM methods like the above to provide additional functionality. The implementation of script header rendering is pretty straight forward – as long as a server header (ie. it has to have runat=”server” set) is available. Otherwise these routines fall back to using the default document level insertions of ScriptManager/ClientScript. Given that there is a server header it’s relatively easy to generate the script tags and code and append them to the header either at the top or bottom. I suspect Microsoft didn’t provide header rendering functionality precisely because a runat=”server” header is not required by ASP.NET so behavior would be slightly unpredictable. That’s not really a problem for a custom implementation however. Here’s the RegisterClientScriptBlock implementation that takes a ScriptRenderModes parameter to allow header rendering: /// <summary> /// Renders client script block with the option of rendering the script block in /// the Html header /// /// For this to work Header must be defined as runat="server" /// </summary> /// <param name="control">any control that instance typically page</param> /// <param name="type">Type that identifies this rendering</param> /// <param name="key">unique script block id</param> /// <param name="script">The script code to render</param> /// <param name="addScriptTags">Ignored for header rendering used for all other insertions</param> /// <param name="renderMode">Where the block is rendered</param> public void RegisterClientScriptBlock(Control control, Type type, string key, string script, bool addScriptTags, ScriptRenderModes renderMode) { if (renderMode == ScriptRenderModes.Inherit) renderMode = DefaultScriptRenderMode; if (control.Page.Header == null || renderMode != ScriptRenderModes.HeaderTop && renderMode != ScriptRenderModes.Header && renderMode != ScriptRenderModes.BottomOfPage) { RegisterClientScriptBlock(control, type, key, script, addScriptTags); return; } // No dupes - ref script include only once const string identifier = "scriptblock_"; if (HttpContext.Current.Items.Contains(identifier + key)) return; HttpContext.Current.Items.Add(identifier + key, string.Empty); StringBuilder sb = new StringBuilder(); // Embed in header sb.AppendLine("\r\n<script type=\"text/javascript\">"); sb.AppendLine(script); sb.AppendLine("</script>"); int? index = HttpContext.Current.Items["__ScriptResourceIndex"] as int?; if (index == null) index = 0; if (renderMode == ScriptRenderModes.HeaderTop) { control.Page.Header.Controls.AddAt(index.Value, new LiteralControl(sb.ToString())); index++; } else if(renderMode == ScriptRenderModes.Header) control.Page.Header.Controls.Add(new LiteralControl(sb.ToString())); else if (renderMode == ScriptRenderModes.BottomOfPage) control.Page.Controls.AddAt(control.Page.Controls.Count-1,new LiteralControl(sb.ToString())); HttpContext.Current.Items["__ScriptResourceIndex"] = index; } Note that the routine has to keep track of items inserted by id so that if the same item is added again with the same key it won’t generate two script entries. Additionally the code has to keep track of how many insertions have been made at the top of the document so that entries are added in the proper order. The RegisterScriptInclude method is similar but there’s some additional logic in here to deal with script file references and ClientScriptProxy’s (optional) custom resource handler that provides script compression /// <summary> /// Registers a client script reference into the page with the option to specify /// the script location in the page /// </summary> /// <param name="control">Any control instance - typically page</param> /// <param name="type">Type that acts as qualifier (uniqueness)</param> /// <param name="url">the Url to the script resource</param> /// <param name="ScriptRenderModes">Determines where the script is rendered</param> public void RegisterClientScriptInclude(Control control, Type type, string url, ScriptRenderModes renderMode) { const string STR_ScriptResourceIndex = "__ScriptResourceIndex"; if (string.IsNullOrEmpty(url)) return; if (renderMode == ScriptRenderModes.Inherit) renderMode = DefaultScriptRenderMode; // Extract just the script filename string fileId = null; // Check resource IDs and try to match to mapped file resources // Used to allow scripts not to be loaded more than once whether // embedded manually (script tag) or via resources with ClientScriptProxy if (url.Contains(".axd?r=")) { string res = HttpUtility.UrlDecode( StringUtils.ExtractString(url, "?r=", "&", false, true) ); foreach (ScriptResourceAlias item in ScriptResourceAliases) { if (item.Resource == res) { fileId = item.Alias + ".js"; break; } } if (fileId == null) fileId = url.ToLower(); } else fileId = Path.GetFileName(url).ToLower(); // No dupes - ref script include only once const string identifier = "script_"; if (HttpContext.Current.Items.Contains( identifier + fileId ) ) return; HttpContext.Current.Items.Add(identifier + fileId, string.Empty); // just use script manager or ClientScriptManager if (control.Page.Header == null || renderMode == ScriptRenderModes.Script || renderMode == ScriptRenderModes.Inline) { RegisterClientScriptInclude(control, type,url, url); return; } // Retrieve script index in header int? index = HttpContext.Current.Items[STR_ScriptResourceIndex] as int?; if (index == null) index = 0; StringBuilder sb = new StringBuilder(256); url = WebUtils.ResolveUrl(url); // Embed in header sb.AppendLine("\r\n<script src=\"" + url + "\" type=\"text/javascript\"></script>"); if (renderMode == ScriptRenderModes.HeaderTop) { control.Page.Header.Controls.AddAt(index.Value, new LiteralControl(sb.ToString())); index++; } else if (renderMode == ScriptRenderModes.Header) control.Page.Header.Controls.Add(new LiteralControl(sb.ToString())); else if (renderMode == ScriptRenderModes.BottomOfPage) control.Page.Controls.AddAt(control.Page.Controls.Count-1, new LiteralControl(sb.ToString())); HttpContext.Current.Items[STR_ScriptResourceIndex] = index; } There’s a little more code here that deals with cleaning up the passed in Url and also some custom handling of script resources that run through the ScriptCompressionModule – any script resources loaded in this fashion are automatically cached based on the resource id. Raw urls extract just the filename from the URL and cache based on that. All of this to avoid doubling up of scripts if called multiple times by multiple instances of the same control for example or several controls that all load the same resources/includes. Finally RegisterClientScriptResource utilizes the previous method to wrap the WebResourceUrl as well as some custom functionality for the resource compression module: /// <summary> /// Returns a WebResource or ScriptResource URL for script resources that are to be /// embedded as script includes. /// </summary> /// <param name="control">Any control</param> /// <param name="type">A type in assembly where resources are located</param> /// <param name="resourceName">Name of the resource to load</param> /// <param name="renderMode">Determines where in the document the link is rendered</param> public void RegisterClientScriptResource(Control control, Type type, string resourceName, ScriptRenderModes renderMode) { string resourceUrl = GetClientScriptResourceUrl(control, type, resourceName); RegisterClientScriptInclude(control, type, resourceUrl, renderMode); } /// <summary> /// Works like GetWebResourceUrl but can be used with javascript resources /// to allow using of resource compression (if the module is loaded). /// </summary> /// <param name="control"></param> /// <param name="type"></param> /// <param name="resourceName"></param> /// <returns></returns> public string GetClientScriptResourceUrl(Control control, Type type, string resourceName) { #if IncludeScriptCompressionModuleSupport // If wwScriptCompression Module through Web.config is loaded use it to compress // script resources by using wcSC.axd Url the module intercepts if (ScriptCompressionModule.ScriptCompressionModuleActive) { string url = "~/wwSC.axd?r=" + HttpUtility.UrlEncode(resourceName); if (type.Assembly != GetType().Assembly) url += "&t=" + HttpUtility.UrlEncode(type.FullName); return WebUtils.ResolveUrl(url); } #endif return control.Page.ClientScript.GetWebResourceUrl(type, resourceName); } This code merely retrieves the resource URL and then simply calls back to RegisterClientScriptInclude with the URL to be embedded which means there’s nothing specific to deal with other than the custom compression module logic which is nice and easy. What else is there in ClientScriptProxy? ClientscriptProxy also provides a few other useful services beyond what I’ve already covered here: Transparent ScriptManager and ClientScript calls ClientScriptProxy includes a host of routines that help figure out whether a script manager is available or not and all functions in this class call the appropriate object – ScriptManager or ClientScript – that is available in the current page to ensure that scripts get embedded into pages properly. This is especially useful for control development where controls have no control over the scripting environment in place on the page. RegisterCssLink and RegisterCssResource Much like the script embedding functions these two methods allow embedding of CSS links. CSS links are appended to the header or to a form declared with runat=”server”. LoadControlScript Is a high level resource loading routine that can be used to easily switch between different script linking modes. It supports loading from a WebResource, a url or not loading anything at all. This is very useful if you build controls that deal with specification of resource urls/ids in a standard way. Check out the full Code You can check out the full code to the ClientScriptProxyClass here: ClientScriptProxy.cs ClientScriptProxy Documentation (class reference) Note that the ClientScriptProxy has a few dependencies in the West Wind Web Toolkit of which it is part of. ControlResources holds a few standard constants and script resource links and the ScriptCompressionModule which is referenced in a few of the script inclusion methods. There’s also another useful ScriptContainer companion control  to the ClientScriptProxy that allows scripts to be placed onto the page’s markup including the ability to specify the script location and script minification options. You can find all the dependencies in the West Wind Web Toolkit repository: West Wind Web Toolkit Repository West Wind Web Toolkit Home Page© Rick Strahl, West Wind Technologies, 2005-2010Posted in ASP.NET  JavaScript  

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  • Improving Partitioned Table Join Performance

    - by Paul White
    The query optimizer does not always choose an optimal strategy when joining partitioned tables. This post looks at an example, showing how a manual rewrite of the query can almost double performance, while reducing the memory grant to almost nothing. Test Data The two tables in this example use a common partitioning partition scheme. The partition function uses 41 equal-size partitions: CREATE PARTITION FUNCTION PFT (integer) AS RANGE RIGHT FOR VALUES ( 125000, 250000, 375000, 500000, 625000, 750000, 875000, 1000000, 1125000, 1250000, 1375000, 1500000, 1625000, 1750000, 1875000, 2000000, 2125000, 2250000, 2375000, 2500000, 2625000, 2750000, 2875000, 3000000, 3125000, 3250000, 3375000, 3500000, 3625000, 3750000, 3875000, 4000000, 4125000, 4250000, 4375000, 4500000, 4625000, 4750000, 4875000, 5000000 ); GO CREATE PARTITION SCHEME PST AS PARTITION PFT ALL TO ([PRIMARY]); There two tables are: CREATE TABLE dbo.T1 ( TID integer NOT NULL IDENTITY(0,1), Column1 integer NOT NULL, Padding binary(100) NOT NULL DEFAULT 0x,   CONSTRAINT PK_T1 PRIMARY KEY CLUSTERED (TID) ON PST (TID) );   CREATE TABLE dbo.T2 ( TID integer NOT NULL, Column1 integer NOT NULL, Padding binary(100) NOT NULL DEFAULT 0x,   CONSTRAINT PK_T2 PRIMARY KEY CLUSTERED (TID, Column1) ON PST (TID) ); The next script loads 5 million rows into T1 with a pseudo-random value between 1 and 5 for Column1. The table is partitioned on the IDENTITY column TID: INSERT dbo.T1 WITH (TABLOCKX) (Column1) SELECT (ABS(CHECKSUM(NEWID())) % 5) + 1 FROM dbo.Numbers AS N WHERE n BETWEEN 1 AND 5000000; In case you don’t already have an auxiliary table of numbers lying around, here’s a script to create one with 10 million rows: CREATE TABLE dbo.Numbers (n bigint PRIMARY KEY);   WITH L0 AS(SELECT 1 AS c UNION ALL SELECT 1), L1 AS(SELECT 1 AS c FROM L0 AS A CROSS JOIN L0 AS B), L2 AS(SELECT 1 AS c FROM L1 AS A CROSS JOIN L1 AS B), L3 AS(SELECT 1 AS c FROM L2 AS A CROSS JOIN L2 AS B), L4 AS(SELECT 1 AS c FROM L3 AS A CROSS JOIN L3 AS B), L5 AS(SELECT 1 AS c FROM L4 AS A CROSS JOIN L4 AS B), Nums AS(SELECT ROW_NUMBER() OVER (ORDER BY (SELECT NULL)) AS n FROM L5) INSERT dbo.Numbers WITH (TABLOCKX) SELECT TOP (10000000) n FROM Nums ORDER BY n OPTION (MAXDOP 1); Table T1 contains data like this: Next we load data into table T2. The relationship between the two tables is that table 2 contains ‘n’ rows for each row in table 1, where ‘n’ is determined by the value in Column1 of table T1. There is nothing particularly special about the data or distribution, by the way. INSERT dbo.T2 WITH (TABLOCKX) (TID, Column1) SELECT T.TID, N.n FROM dbo.T1 AS T JOIN dbo.Numbers AS N ON N.n >= 1 AND N.n <= T.Column1; Table T2 ends up containing about 15 million rows: The primary key for table T2 is a combination of TID and Column1. The data is partitioned according to the value in column TID alone. Partition Distribution The following query shows the number of rows in each partition of table T1: SELECT PartitionID = CA1.P, NumRows = COUNT_BIG(*) FROM dbo.T1 AS T CROSS APPLY (VALUES ($PARTITION.PFT(TID))) AS CA1 (P) GROUP BY CA1.P ORDER BY CA1.P; There are 40 partitions containing 125,000 rows (40 * 125k = 5m rows). The rightmost partition remains empty. The next query shows the distribution for table 2: SELECT PartitionID = CA1.P, NumRows = COUNT_BIG(*) FROM dbo.T2 AS T CROSS APPLY (VALUES ($PARTITION.PFT(TID))) AS CA1 (P) GROUP BY CA1.P ORDER BY CA1.P; There are roughly 375,000 rows in each partition (the rightmost partition is also empty): Ok, that’s the test data done. Test Query and Execution Plan The task is to count the rows resulting from joining tables 1 and 2 on the TID column: SET STATISTICS IO ON; DECLARE @s datetime2 = SYSUTCDATETIME();   SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID;   SELECT DATEDIFF(Millisecond, @s, SYSUTCDATETIME()); SET STATISTICS IO OFF; The optimizer chooses a plan using parallel hash join, and partial aggregation: The Plan Explorer plan tree view shows accurate cardinality estimates and an even distribution of rows across threads (click to enlarge the image): With a warm data cache, the STATISTICS IO output shows that no physical I/O was needed, and all 41 partitions were touched: Running the query without actual execution plan or STATISTICS IO information for maximum performance, the query returns in around 2600ms. Execution Plan Analysis The first step toward improving on the execution plan produced by the query optimizer is to understand how it works, at least in outline. The two parallel Clustered Index Scans use multiple threads to read rows from tables T1 and T2. Parallel scan uses a demand-based scheme where threads are given page(s) to scan from the table as needed. This arrangement has certain important advantages, but does result in an unpredictable distribution of rows amongst threads. The point is that multiple threads cooperate to scan the whole table, but it is impossible to predict which rows end up on which threads. For correct results from the parallel hash join, the execution plan has to ensure that rows from T1 and T2 that might join are processed on the same thread. For example, if a row from T1 with join key value ‘1234’ is placed in thread 5’s hash table, the execution plan must guarantee that any rows from T2 that also have join key value ‘1234’ probe thread 5’s hash table for matches. The way this guarantee is enforced in this parallel hash join plan is by repartitioning rows to threads after each parallel scan. The two repartitioning exchanges route rows to threads using a hash function over the hash join keys. The two repartitioning exchanges use the same hash function so rows from T1 and T2 with the same join key must end up on the same hash join thread. Expensive Exchanges This business of repartitioning rows between threads can be very expensive, especially if a large number of rows is involved. The execution plan selected by the optimizer moves 5 million rows through one repartitioning exchange and around 15 million across the other. As a first step toward removing these exchanges, consider the execution plan selected by the optimizer if we join just one partition from each table, disallowing parallelism: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = 1 AND $PARTITION.PFT(T2.TID) = 1 OPTION (MAXDOP 1); The optimizer has chosen a (one-to-many) merge join instead of a hash join. The single-partition query completes in around 100ms. If everything scaled linearly, we would expect that extending this strategy to all 40 populated partitions would result in an execution time around 4000ms. Using parallelism could reduce that further, perhaps to be competitive with the parallel hash join chosen by the optimizer. This raises a question. If the most efficient way to join one partition from each of the tables is to use a merge join, why does the optimizer not choose a merge join for the full query? Forcing a Merge Join Let’s force the optimizer to use a merge join on the test query using a hint: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (MERGE JOIN); This is the execution plan selected by the optimizer: This plan results in the same number of logical reads reported previously, but instead of 2600ms the query takes 5000ms. The natural explanation for this drop in performance is that the merge join plan is only using a single thread, whereas the parallel hash join plan could use multiple threads. Parallel Merge Join We can get a parallel merge join plan using the same query hint as before, and adding trace flag 8649: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (MERGE JOIN, QUERYTRACEON 8649); The execution plan is: This looks promising. It uses a similar strategy to distribute work across threads as seen for the parallel hash join. In practice though, performance is disappointing. On a typical run, the parallel merge plan runs for around 8400ms; slower than the single-threaded merge join plan (5000ms) and much worse than the 2600ms for the parallel hash join. We seem to be going backwards! The logical reads for the parallel merge are still exactly the same as before, with no physical IOs. The cardinality estimates and thread distribution are also still very good (click to enlarge): A big clue to the reason for the poor performance is shown in the wait statistics (captured by Plan Explorer Pro): CXPACKET waits require careful interpretation, and are most often benign, but in this case excessive waiting occurs at the repartitioning exchanges. Unlike the parallel hash join, the repartitioning exchanges in this plan are order-preserving ‘merging’ exchanges (because merge join requires ordered inputs): Parallelism works best when threads can just grab any available unit of work and get on with processing it. Preserving order introduces inter-thread dependencies that can easily lead to significant waits occurring. In extreme cases, these dependencies can result in an intra-query deadlock, though the details of that will have to wait for another time to explore in detail. The potential for waits and deadlocks leads the query optimizer to cost parallel merge join relatively highly, especially as the degree of parallelism (DOP) increases. This high costing resulted in the optimizer choosing a serial merge join rather than parallel in this case. The test results certainly confirm its reasoning. Collocated Joins In SQL Server 2008 and later, the optimizer has another available strategy when joining tables that share a common partition scheme. This strategy is a collocated join, also known as as a per-partition join. It can be applied in both serial and parallel execution plans, though it is limited to 2-way joins in the current optimizer. Whether the optimizer chooses a collocated join or not depends on cost estimation. The primary benefits of a collocated join are that it eliminates an exchange and requires less memory, as we will see next. Costing and Plan Selection The query optimizer did consider a collocated join for our original query, but it was rejected on cost grounds. The parallel hash join with repartitioning exchanges appeared to be a cheaper option. There is no query hint to force a collocated join, so we have to mess with the costing framework to produce one for our test query. Pretending that IOs cost 50 times more than usual is enough to convince the optimizer to use collocated join with our test query: -- Pretend IOs are 50x cost temporarily DBCC SETIOWEIGHT(50);   -- Co-located hash join SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (RECOMPILE);   -- Reset IO costing DBCC SETIOWEIGHT(1); Collocated Join Plan The estimated execution plan for the collocated join is: The Constant Scan contains one row for each partition of the shared partitioning scheme, from 1 to 41. The hash repartitioning exchanges seen previously are replaced by a single Distribute Streams exchange using Demand partitioning. Demand partitioning means that the next partition id is given to the next parallel thread that asks for one. My test machine has eight logical processors, and all are available for SQL Server to use. As a result, there are eight threads in the single parallel branch in this plan, each processing one partition from each table at a time. Once a thread finishes processing a partition, it grabs a new partition number from the Distribute Streams exchange…and so on until all partitions have been processed. It is important to understand that the parallel scans in this plan are different from the parallel hash join plan. Although the scans have the same parallelism icon, tables T1 and T2 are not being co-operatively scanned by multiple threads in the same way. Each thread reads a single partition of T1 and performs a hash match join with the same partition from table T2. The properties of the two Clustered Index Scans show a Seek Predicate (unusual for a scan!) limiting the rows to a single partition: The crucial point is that the join between T1 and T2 is on TID, and TID is the partitioning column for both tables. A thread that processes partition ‘n’ is guaranteed to see all rows that can possibly join on TID for that partition. In addition, no other thread will see rows from that partition, so this removes the need for repartitioning exchanges. CPU and Memory Efficiency Improvements The collocated join has removed two expensive repartitioning exchanges and added a single exchange processing 41 rows (one for each partition id). Remember, the parallel hash join plan exchanges had to process 5 million and 15 million rows. The amount of processor time spent on exchanges will be much lower in the collocated join plan. In addition, the collocated join plan has a maximum of 8 threads processing single partitions at any one time. The 41 partitions will all be processed eventually, but a new partition is not started until a thread asks for it. Threads can reuse hash table memory for the new partition. The parallel hash join plan also had 8 hash tables, but with all 5,000,000 build rows loaded at the same time. The collocated plan needs memory for only 8 * 125,000 = 1,000,000 rows at any one time. Collocated Hash Join Performance The collated join plan has disappointing performance in this case. The query runs for around 25,300ms despite the same IO statistics as usual. This is much the worst result so far, so what went wrong? It turns out that cardinality estimation for the single partition scans of table T1 is slightly low. The properties of the Clustered Index Scan of T1 (graphic immediately above) show the estimation was for 121,951 rows. This is a small shortfall compared with the 125,000 rows actually encountered, but it was enough to cause the hash join to spill to physical tempdb: A level 1 spill doesn’t sound too bad, until you realize that the spill to tempdb probably occurs for each of the 41 partitions. As a side note, the cardinality estimation error is a little surprising because the system tables accurately show there are 125,000 rows in every partition of T1. Unfortunately, the optimizer uses regular column and index statistics to derive cardinality estimates here rather than system table information (e.g. sys.partitions). Collocated Merge Join We will never know how well the collocated parallel hash join plan might have worked without the cardinality estimation error (and the resulting 41 spills to tempdb) but we do know: Merge join does not require a memory grant; and Merge join was the optimizer’s preferred join option for a single partition join Putting this all together, what we would really like to see is the same collocated join strategy, but using merge join instead of hash join. Unfortunately, the current query optimizer cannot produce a collocated merge join; it only knows how to do collocated hash join. So where does this leave us? CROSS APPLY sys.partitions We can try to write our own collocated join query. We can use sys.partitions to find the partition numbers, and CROSS APPLY to get a count per partition, with a final step to sum the partial counts. The following query implements this idea: SELECT row_count = SUM(Subtotals.cnt) FROM ( -- Partition numbers SELECT p.partition_number FROM sys.partitions AS p WHERE p.[object_id] = OBJECT_ID(N'T1', N'U') AND p.index_id = 1 ) AS P CROSS APPLY ( -- Count per collocated join SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals; The estimated plan is: The cardinality estimates aren’t all that good here, especially the estimate for the scan of the system table underlying the sys.partitions view. Nevertheless, the plan shape is heading toward where we would like to be. Each partition number from the system table results in a per-partition scan of T1 and T2, a one-to-many Merge Join, and a Stream Aggregate to compute the partial counts. The final Stream Aggregate just sums the partial counts. Execution time for this query is around 3,500ms, with the same IO statistics as always. This compares favourably with 5,000ms for the serial plan produced by the optimizer with the OPTION (MERGE JOIN) hint. This is another case of the sum of the parts being less than the whole – summing 41 partial counts from 41 single-partition merge joins is faster than a single merge join and count over all partitions. Even so, this single-threaded collocated merge join is not as quick as the original parallel hash join plan, which executed in 2,600ms. On the positive side, our collocated merge join uses only one logical processor and requires no memory grant. The parallel hash join plan used 16 threads and reserved 569 MB of memory:   Using a Temporary Table Our collocated merge join plan should benefit from parallelism. The reason parallelism is not being used is that the query references a system table. We can work around that by writing the partition numbers to a temporary table (or table variable): SET STATISTICS IO ON; DECLARE @s datetime2 = SYSUTCDATETIME();   CREATE TABLE #P ( partition_number integer PRIMARY KEY);   INSERT #P (partition_number) SELECT p.partition_number FROM sys.partitions AS p WHERE p.[object_id] = OBJECT_ID(N'T1', N'U') AND p.index_id = 1;   SELECT row_count = SUM(Subtotals.cnt) FROM #P AS p CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals;   DROP TABLE #P;   SELECT DATEDIFF(Millisecond, @s, SYSUTCDATETIME()); SET STATISTICS IO OFF; Using the temporary table adds a few logical reads, but the overall execution time is still around 3500ms, indistinguishable from the same query without the temporary table. The problem is that the query optimizer still doesn’t choose a parallel plan for this query, though the removal of the system table reference means that it could if it chose to: In fact the optimizer did enter the parallel plan phase of query optimization (running search 1 for a second time): Unfortunately, the parallel plan found seemed to be more expensive than the serial plan. This is a crazy result, caused by the optimizer’s cost model not reducing operator CPU costs on the inner side of a nested loops join. Don’t get me started on that, we’ll be here all night. In this plan, everything expensive happens on the inner side of a nested loops join. Without a CPU cost reduction to compensate for the added cost of exchange operators, candidate parallel plans always look more expensive to the optimizer than the equivalent serial plan. Parallel Collocated Merge Join We can produce the desired parallel plan using trace flag 8649 again: SELECT row_count = SUM(Subtotals.cnt) FROM #P AS p CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals OPTION (QUERYTRACEON 8649); The actual execution plan is: One difference between this plan and the collocated hash join plan is that a Repartition Streams exchange operator is used instead of Distribute Streams. The effect is similar, though not quite identical. The Repartition uses round-robin partitioning, meaning the next partition id is pushed to the next thread in sequence. The Distribute Streams exchange seen earlier used Demand partitioning, meaning the next partition id is pulled across the exchange by the next thread that is ready for more work. There are subtle performance implications for each partitioning option, but going into that would again take us too far off the main point of this post. Performance The important thing is the performance of this parallel collocated merge join – just 1350ms on a typical run. The list below shows all the alternatives from this post (all timings include creation, population, and deletion of the temporary table where appropriate) from quickest to slowest: Collocated parallel merge join: 1350ms Parallel hash join: 2600ms Collocated serial merge join: 3500ms Serial merge join: 5000ms Parallel merge join: 8400ms Collated parallel hash join: 25,300ms (hash spill per partition) The parallel collocated merge join requires no memory grant (aside from a paltry 1.2MB used for exchange buffers). This plan uses 16 threads at DOP 8; but 8 of those are (rather pointlessly) allocated to the parallel scan of the temporary table. These are minor concerns, but it turns out there is a way to address them if it bothers you. Parallel Collocated Merge Join with Demand Partitioning This final tweak replaces the temporary table with a hard-coded list of partition ids (dynamic SQL could be used to generate this query from sys.partitions): SELECT row_count = SUM(Subtotals.cnt) FROM ( VALUES (1),(2),(3),(4),(5),(6),(7),(8),(9),(10), (11),(12),(13),(14),(15),(16),(17),(18),(19),(20), (21),(22),(23),(24),(25),(26),(27),(28),(29),(30), (31),(32),(33),(34),(35),(36),(37),(38),(39),(40),(41) ) AS P (partition_number) CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals OPTION (QUERYTRACEON 8649); The actual execution plan is: The parallel collocated hash join plan is reproduced below for comparison: The manual rewrite has another advantage that has not been mentioned so far: the partial counts (per partition) can be computed earlier than the partial counts (per thread) in the optimizer’s collocated join plan. The earlier aggregation is performed by the extra Stream Aggregate under the nested loops join. The performance of the parallel collocated merge join is unchanged at around 1350ms. Final Words It is a shame that the current query optimizer does not consider a collocated merge join (Connect item closed as Won’t Fix). The example used in this post showed an improvement in execution time from 2600ms to 1350ms using a modestly-sized data set and limited parallelism. In addition, the memory requirement for the query was almost completely eliminated  – down from 569MB to 1.2MB. The problem with the parallel hash join selected by the optimizer is that it attempts to process the full data set all at once (albeit using eight threads). It requires a large memory grant to hold all 5 million rows from table T1 across the eight hash tables, and does not take advantage of the divide-and-conquer opportunity offered by the common partitioning. The great thing about the collocated join strategies is that each parallel thread works on a single partition from both tables, reading rows, performing the join, and computing a per-partition subtotal, before moving on to a new partition. From a thread’s point of view… If you have trouble visualizing what is happening from just looking at the parallel collocated merge join execution plan, let’s look at it again, but from the point of view of just one thread operating between the two Parallelism (exchange) operators. Our thread picks up a single partition id from the Distribute Streams exchange, and starts a merge join using ordered rows from partition 1 of table T1 and partition 1 of table T2. By definition, this is all happening on a single thread. As rows join, they are added to a (per-partition) count in the Stream Aggregate immediately above the Merge Join. Eventually, either T1 (partition 1) or T2 (partition 1) runs out of rows and the merge join stops. The per-partition count from the aggregate passes on through the Nested Loops join to another Stream Aggregate, which is maintaining a per-thread subtotal. Our same thread now picks up a new partition id from the exchange (say it gets id 9 this time). The count in the per-partition aggregate is reset to zero, and the processing of partition 9 of both tables proceeds just as it did for partition 1, and on the same thread. Each thread picks up a single partition id and processes all the data for that partition, completely independently from other threads working on other partitions. One thread might eventually process partitions (1, 9, 17, 25, 33, 41) while another is concurrently processing partitions (2, 10, 18, 26, 34) and so on for the other six threads at DOP 8. The point is that all 8 threads can execute independently and concurrently, continuing to process new partitions until the wider job (of which the thread has no knowledge!) is done. This divide-and-conquer technique can be much more efficient than simply splitting the entire workload across eight threads all at once. Related Reading Understanding and Using Parallelism in SQL Server Parallel Execution Plans Suck © 2013 Paul White – All Rights Reserved Twitter: @SQL_Kiwi

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  • Toorcon 15 (2013)

    - by danx
    The Toorcon gang (senior staff): h1kari (founder), nfiltr8, and Geo Introduction to Toorcon 15 (2013) A Tale of One Software Bypass of MS Windows 8 Secure Boot Breaching SSL, One Byte at a Time Running at 99%: Surviving an Application DoS Security Response in the Age of Mass Customized Attacks x86 Rewriting: Defeating RoP and other Shinanighans Clowntown Express: interesting bugs and running a bug bounty program Active Fingerprinting of Encrypted VPNs Making Attacks Go Backwards Mask Your Checksums—The Gorry Details Adventures with weird machines thirty years after "Reflections on Trusting Trust" Introduction to Toorcon 15 (2013) Toorcon 15 is the 15th annual security conference held in San Diego. I've attended about a third of them and blogged about previous conferences I attended here starting in 2003. As always, I've only summarized the talks I attended and interested me enough to write about them. Be aware that I may have misrepresented the speaker's remarks and that they are not my remarks or opinion, or those of my employer, so don't quote me or them. Those seeking further details may contact the speakers directly or use The Google. For some talks, I have a URL for further information. A Tale of One Software Bypass of MS Windows 8 Secure Boot Andrew Furtak and Oleksandr Bazhaniuk Yuri Bulygin, Oleksandr ("Alex") Bazhaniuk, and (not present) Andrew Furtak Yuri and Alex talked about UEFI and Bootkits and bypassing MS Windows 8 Secure Boot, with vendor recommendations. They previously gave this talk at the BlackHat 2013 conference. MS Windows 8 Secure Boot Overview UEFI (Unified Extensible Firmware Interface) is interface between hardware and OS. UEFI is processor and architecture independent. Malware can replace bootloader (bootx64.efi, bootmgfw.efi). Once replaced can modify kernel. Trivial to replace bootloader. Today many legacy bootkits—UEFI replaces them most of them. MS Windows 8 Secure Boot verifies everything you load, either through signatures or hashes. UEFI firmware relies on secure update (with signed update). You would think Secure Boot would rely on ROM (such as used for phones0, but you can't do that for PCs—PCs use writable memory with signatures DXE core verifies the UEFI boat loader(s) OS Loader (winload.efi, winresume.efi) verifies the OS kernel A chain of trust is established with a root key (Platform Key, PK), which is a cert belonging to the platform vendor. Key Exchange Keys (KEKs) verify an "authorized" database (db), and "forbidden" database (dbx). X.509 certs with SHA-1/SHA-256 hashes. Keys are stored in non-volatile (NV) flash-based NVRAM. Boot Services (BS) allow adding/deleting keys (can't be accessed once OS starts—which uses Run-Time (RT)). Root cert uses RSA-2048 public keys and PKCS#7 format signatures. SecureBoot — enable disable image signature checks SetupMode — update keys, self-signed keys, and secure boot variables CustomMode — allows updating keys Secure Boot policy settings are: always execute, never execute, allow execute on security violation, defer execute on security violation, deny execute on security violation, query user on security violation Attacking MS Windows 8 Secure Boot Secure Boot does NOT protect from physical access. Can disable from console. Each BIOS vendor implements Secure Boot differently. There are several platform and BIOS vendors. It becomes a "zoo" of implementations—which can be taken advantage of. Secure Boot is secure only when all vendors implement it correctly. Allow only UEFI firmware signed updates protect UEFI firmware from direct modification in flash memory protect FW update components program SPI controller securely protect secure boot policy settings in nvram protect runtime api disable compatibility support module which allows unsigned legacy Can corrupt the Platform Key (PK) EFI root certificate variable in SPI flash. If PK is not found, FW enters setup mode wich secure boot turned off. Can also exploit TPM in a similar manner. One is not supposed to be able to directly modify the PK in SPI flash from the OS though. But they found a bug that they can exploit from User Mode (undisclosed) and demoed the exploit. It loaded and ran their own bootkit. The exploit requires a reboot. Multiple vendors are vulnerable. They will disclose this exploit to vendors in the future. Recommendations: allow only signed updates protect UEFI fw in ROM protect EFI variable store in ROM Breaching SSL, One Byte at a Time Yoel Gluck and Angelo Prado Angelo Prado and Yoel Gluck, Salesforce.com CRIME is software that performs a "compression oracle attack." This is possible because the SSL protocol doesn't hide length, and because SSL compresses the header. CRIME requests with every possible character and measures the ciphertext length. Look for the plaintext which compresses the most and looks for the cookie one byte-at-a-time. SSL Compression uses LZ77 to reduce redundancy. Huffman coding replaces common byte sequences with shorter codes. US CERT thinks the SSL compression problem is fixed, but it isn't. They convinced CERT that it wasn't fixed and they issued a CVE. BREACH, breachattrack.com BREACH exploits the SSL response body (Accept-Encoding response, Content-Encoding). It takes advantage of the fact that the response is not compressed. BREACH uses gzip and needs fairly "stable" pages that are static for ~30 seconds. It needs attacker-supplied content (say from a web form or added to a URL parameter). BREACH listens to a session's requests and responses, then inserts extra requests and responses. Eventually, BREACH guesses a session's secret key. Can use compression to guess contents one byte at-a-time. For example, "Supersecret SupersecreX" (a wrong guess) compresses 10 bytes, and "Supersecret Supersecret" (a correct guess) compresses 11 bytes, so it can find each character by guessing every character. To start the guess, BREACH needs at least three known initial characters in the response sequence. Compression length then "leaks" information. Some roadblocks include no winners (all guesses wrong) or too many winners (multiple possibilities that compress the same). The solutions include: lookahead (guess 2 or 3 characters at-a-time instead of 1 character). Expensive rollback to last known conflict check compression ratio can brute-force first 3 "bootstrap" characters, if needed (expensive) block ciphers hide exact plain text length. Solution is to align response in advance to block size Mitigations length: use variable padding secrets: dynamic CSRF tokens per request secret: change over time separate secret to input-less servlets Future work eiter understand DEFLATE/GZIP HTTPS extensions Running at 99%: Surviving an Application DoS Ryan Huber Ryan Huber, Risk I/O Ryan first discussed various ways to do a denial of service (DoS) attack against web services. One usual method is to find a slow web page and do several wgets. Or download large files. Apache is not well suited at handling a large number of connections, but one can put something in front of it Can use Apache alternatives, such as nginx How to identify malicious hosts short, sudden web requests user-agent is obvious (curl, python) same url requested repeatedly no web page referer (not normal) hidden links. hide a link and see if a bot gets it restricted access if not your geo IP (unless the website is global) missing common headers in request regular timing first seen IP at beginning of attack count requests per hosts (usually a very large number) Use of captcha can mitigate attacks, but you'll lose a lot of genuine users. Bouncer, goo.gl/c2vyEc and www.github.com/rawdigits/Bouncer Bouncer is software written by Ryan in netflow. Bouncer has a small, unobtrusive footprint and detects DoS attempts. It closes blacklisted sockets immediately (not nice about it, no proper close connection). Aggregator collects requests and controls your web proxies. Need NTP on the front end web servers for clean data for use by bouncer. Bouncer is also useful for a popularity storm ("Slashdotting") and scraper storms. Future features: gzip collection data, documentation, consumer library, multitask, logging destroyed connections. Takeaways: DoS mitigation is easier with a complete picture Bouncer designed to make it easier to detect and defend DoS—not a complete cure Security Response in the Age of Mass Customized Attacks Peleus Uhley and Karthik Raman Peleus Uhley and Karthik Raman, Adobe ASSET, blogs.adobe.com/asset/ Peleus and Karthik talked about response to mass-customized exploits. Attackers behave much like a business. "Mass customization" refers to concept discussed in the book Future Perfect by Stan Davis of Harvard Business School. Mass customization is differentiating a product for an individual customer, but at a mass production price. For example, the same individual with a debit card receives basically the same customized ATM experience around the world. Or designing your own PC from commodity parts. Exploit kits are another example of mass customization. The kits support multiple browsers and plugins, allows new modules. Exploit kits are cheap and customizable. Organized gangs use exploit kits. A group at Berkeley looked at 77,000 malicious websites (Grier et al., "Manufacturing Compromise: The Emergence of Exploit-as-a-Service", 2012). They found 10,000 distinct binaries among them, but derived from only a dozen or so exploit kits. Characteristics of Mass Malware: potent, resilient, relatively low cost Technical characteristics: multiple OS, multipe payloads, multiple scenarios, multiple languages, obfuscation Response time for 0-day exploits has gone down from ~40 days 5 years ago to about ~10 days now. So the drive with malware is towards mass customized exploits, to avoid detection There's plenty of evicence that exploit development has Project Manager bureaucracy. They infer from the malware edicts to: support all versions of reader support all versions of windows support all versions of flash support all browsers write large complex, difficult to main code (8750 lines of JavaScript for example Exploits have "loose coupling" of multipe versions of software (adobe), OS, and browser. This allows specific attacks against specific versions of multiple pieces of software. Also allows exploits of more obscure software/OS/browsers and obscure versions. Gave examples of exploits that exploited 2, 3, 6, or 14 separate bugs. However, these complete exploits are more likely to be buggy or fragile in themselves and easier to defeat. Future research includes normalizing malware and Javascript. Conclusion: The coming trend is that mass-malware with mass zero-day attacks will result in mass customization of attacks. x86 Rewriting: Defeating RoP and other Shinanighans Richard Wartell Richard Wartell The attack vector we are addressing here is: First some malware causes a buffer overflow. The malware has no program access, but input access and buffer overflow code onto stack Later the stack became non-executable. The workaround malware used was to write a bogus return address to the stack jumping to malware Later came ASLR (Address Space Layout Randomization) to randomize memory layout and make addresses non-deterministic. The workaround malware used was to jump t existing code segments in the program that can be used in bad ways "RoP" is Return-oriented Programming attacks. RoP attacks use your own code and write return address on stack to (existing) expoitable code found in program ("gadgets"). Pinkie Pie was paid $60K last year for a RoP attack. One solution is using anti-RoP compilers that compile source code with NO return instructions. ASLR does not randomize address space, just "gadgets". IPR/ILR ("Instruction Location Randomization") randomizes each instruction with a virtual machine. Richard's goal was to randomize a binary with no source code access. He created "STIR" (Self-Transofrming Instruction Relocation). STIR disassembles binary and operates on "basic blocks" of code. The STIR disassembler is conservative in what to disassemble. Each basic block is moved to a random location in memory. Next, STIR writes new code sections with copies of "basic blocks" of code in randomized locations. The old code is copied and rewritten with jumps to new code. the original code sections in the file is marked non-executible. STIR has better entropy than ASLR in location of code. Makes brute force attacks much harder. STIR runs on MS Windows (PEM) and Linux (ELF). It eliminated 99.96% or more "gadgets" (i.e., moved the address). Overhead usually 5-10% on MS Windows, about 1.5-4% on Linux (but some code actually runs faster!). The unique thing about STIR is it requires no source access and the modified binary fully works! Current work is to rewrite code to enforce security policies. For example, don't create a *.{exe,msi,bat} file. Or don't connect to the network after reading from the disk. Clowntown Express: interesting bugs and running a bug bounty program Collin Greene Collin Greene, Facebook Collin talked about Facebook's bug bounty program. Background at FB: FB has good security frameworks, such as security teams, external audits, and cc'ing on diffs. But there's lots of "deep, dark, forgotten" parts of legacy FB code. Collin gave several examples of bountied bugs. Some bounty submissions were on software purchased from a third-party (but bounty claimers don't know and don't care). We use security questions, as does everyone else, but they are basically insecure (often easily discoverable). Collin didn't expect many bugs from the bounty program, but they ended getting 20+ good bugs in first 24 hours and good submissions continue to come in. Bug bounties bring people in with different perspectives, and are paid only for success. Bug bounty is a better use of a fixed amount of time and money versus just code review or static code analysis. The Bounty program started July 2011 and paid out $1.5 million to date. 14% of the submissions have been high priority problems that needed to be fixed immediately. The best bugs come from a small % of submitters (as with everything else)—the top paid submitters are paid 6 figures a year. Spammers like to backstab competitors. The youngest sumitter was 13. Some submitters have been hired. Bug bounties also allows to see bugs that were missed by tools or reviews, allowing improvement in the process. Bug bounties might not work for traditional software companies where the product has release cycle or is not on Internet. Active Fingerprinting of Encrypted VPNs Anna Shubina Anna Shubina, Dartmouth Institute for Security, Technology, and Society (I missed the start of her talk because another track went overtime. But I have the DVD of the talk, so I'll expand later) IPsec leaves fingerprints. Using netcat, one can easily visually distinguish various crypto chaining modes just from packet timing on a chart (example, DES-CBC versus AES-CBC) One can tell a lot about VPNs just from ping roundtrips (such as what router is used) Delayed packets are not informative about a network, especially if far away from the network More needed to explore about how TCP works in real life with respect to timing Making Attacks Go Backwards Fuzzynop FuzzyNop, Mandiant This talk is not about threat attribution (finding who), product solutions, politics, or sales pitches. But who are making these malware threats? It's not a single person or group—they have diverse skill levels. There's a lot of fat-fingered fumblers out there. Always look for low-hanging fruit first: "hiding" malware in the temp, recycle, or root directories creation of unnamed scheduled tasks obvious names of files and syscalls ("ClearEventLog") uncleared event logs. Clearing event log in itself, and time of clearing, is a red flag and good first clue to look for on a suspect system Reverse engineering is hard. Disassembler use takes practice and skill. A popular tool is IDA Pro, but it takes multiple interactive iterations to get a clean disassembly. Key loggers are used a lot in targeted attacks. They are typically custom code or built in a backdoor. A big tip-off is that non-printable characters need to be printed out (such as "[Ctrl]" "[RightShift]") or time stamp printf strings. Look for these in files. Presence is not proof they are used. Absence is not proof they are not used. Java exploits. Can parse jar file with idxparser.py and decomile Java file. Java typially used to target tech companies. Backdoors are the main persistence mechanism (provided externally) for malware. Also malware typically needs command and control. Application of Artificial Intelligence in Ad-Hoc Static Code Analysis John Ashaman John Ashaman, Security Innovation Initially John tried to analyze open source files with open source static analysis tools, but these showed thousands of false positives. Also tried using grep, but tis fails to find anything even mildly complex. So next John decided to write his own tool. His approach was to first generate a call graph then analyze the graph. However, the problem is that making a call graph is really hard. For example, one problem is "evil" coding techniques, such as passing function pointer. First the tool generated an Abstract Syntax Tree (AST) with the nodes created from method declarations and edges created from method use. Then the tool generated a control flow graph with the goal to find a path through the AST (a maze) from source to sink. The algorithm is to look at adjacent nodes to see if any are "scary" (a vulnerability), using heuristics for search order. The tool, called "Scat" (Static Code Analysis Tool), currently looks for C# vulnerabilities and some simple PHP. Later, he plans to add more PHP, then JSP and Java. For more information see his posts in Security Innovation blog and NRefactory on GitHub. Mask Your Checksums—The Gorry Details Eric (XlogicX) Davisson Eric (XlogicX) Davisson Sometimes in emailing or posting TCP/IP packets to analyze problems, you may want to mask the IP address. But to do this correctly, you need to mask the checksum too, or you'll leak information about the IP. Problem reports found in stackoverflow.com, sans.org, and pastebin.org are usually not masked, but a few companies do care. If only the IP is masked, the IP may be guessed from checksum (that is, it leaks data). Other parts of packet may leak more data about the IP. TCP and IP checksums both refer to the same data, so can get more bits of information out of using both checksums than just using one checksum. Also, one can usually determine the OS from the TTL field and ports in a packet header. If we get hundreds of possible results (16x each masked nibble that is unknown), one can do other things to narrow the results, such as look at packet contents for domain or geo information. With hundreds of results, can import as CSV format into a spreadsheet. Can corelate with geo data and see where each possibility is located. Eric then demoed a real email report with a masked IP packet attached. Was able to find the exact IP address, given the geo and university of the sender. Point is if you're going to mask a packet, do it right. Eric wouldn't usually bother, but do it correctly if at all, to not create a false impression of security. Adventures with weird machines thirty years after "Reflections on Trusting Trust" Sergey Bratus Sergey Bratus, Dartmouth College (and Julian Bangert and Rebecca Shapiro, not present) "Reflections on Trusting Trust" refers to Ken Thompson's classic 1984 paper. "You can't trust code that you did not totally create yourself." There's invisible links in the chain-of-trust, such as "well-installed microcode bugs" or in the compiler, and other planted bugs. Thompson showed how a compiler can introduce and propagate bugs in unmodified source. But suppose if there's no bugs and you trust the author, can you trust the code? Hell No! There's too many factors—it's Babylonian in nature. Why not? Well, Input is not well-defined/recognized (code's assumptions about "checked" input will be violated (bug/vunerabiliy). For example, HTML is recursive, but Regex checking is not recursive. Input well-formed but so complex there's no telling what it does For example, ELF file parsing is complex and has multiple ways of parsing. Input is seen differently by different pieces of program or toolchain Any Input is a program input executes on input handlers (drives state changes & transitions) only a well-defined execution model can be trusted (regex/DFA, PDA, CFG) Input handler either is a "recognizer" for the inputs as a well-defined language (see langsec.org) or it's a "virtual machine" for inputs to drive into pwn-age ELF ABI (UNIX/Linux executible file format) case study. Problems can arise from these steps (without planting bugs): compiler linker loader ld.so/rtld relocator DWARF (debugger info) exceptions The problem is you can't really automatically analyze code (it's the "halting problem" and undecidable). Only solution is to freeze code and sign it. But you can't freeze everything! Can't freeze ASLR or loading—must have tables and metadata. Any sufficiently complex input data is the same as VM byte code Example, ELF relocation entries + dynamic symbols == a Turing Complete Machine (TM). @bxsays created a Turing machine in Linux from relocation data (not code) in an ELF file. For more information, see Rebecca "bx" Shapiro's presentation from last year's Toorcon, "Programming Weird Machines with ELF Metadata" @bxsays did same thing with Mach-O bytecode Or a DWARF exception handling data .eh_frame + glibc == Turning Machine X86 MMU (IDT, GDT, TSS): used address translation to create a Turning Machine. Page handler reads and writes (on page fault) memory. Uses a page table, which can be used as Turning Machine byte code. Example on Github using this TM that will fly a glider across the screen Next Sergey talked about "Parser Differentials". That having one input format, but two parsers, will create confusion and opportunity for exploitation. For example, CSRs are parsed during creation by cert requestor and again by another parser at the CA. Another example is ELF—several parsers in OS tool chain, which are all different. Can have two different Program Headers (PHDRs) because ld.so parses multiple PHDRs. The second PHDR can completely transform the executable. This is described in paper in the first issue of International Journal of PoC. Conclusions trusting computers not only about bugs! Bugs are part of a problem, but no by far all of it complex data formats means bugs no "chain of trust" in Babylon! (that is, with parser differentials) we need to squeeze complexity out of data until data stops being "code equivalent" Further information See and langsec.org. USENIX WOOT 2013 (Workshop on Offensive Technologies) for "weird machines" papers and videos.

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  • AuthnRequest Settings in OIF / SP

    - by Damien Carru
    In this article, I will list the various OIF/SP settings that affect how an AuthnRequest message is created in OIF in a Federation SSO flow. The AuthnRequest message is used by an SP to start a Federation SSO operation and to indicate to the IdP how the operation should be executed: How the user should be challenged at the IdP Whether or not the user should be challenged at the IdP, even if a session already exists at the IdP for this user Which NameID format should be requested in the SAML Assertion Which binding (Artifact or HTTP-POST) should be requested from the IdP to send the Assertion Which profile should be used by OIF/SP to send the AuthnRequest message Enjoy the reading! Protocols The SAML 2.0, SAML 1.1 and OpenID 2.0 protocols define different message elements and rules that allow an administrator to influence the Federation SSO flows in different manners, when the SP triggers an SSO operation: SAML 2.0 allows extensive customization via the AuthnRequest message SAML 1.1 does not allow any customization, since the specifications do not define an authentication request message OpenID 2.0 allows for some customization, mainly via the OpenID 2.0 extensions such as PAPE or UI SAML 2.0 OIF/SP allows the customization of the SAML 2.0 AuthnRequest message for the following elements: ForceAuthn: Boolean indicating whether or not the IdP should force the user for re-authentication, even if the user has still a valid session By default set to false IsPassive Boolean indicating whether or not the IdP is allowed to interact with the user as part of the Federation SSO operation. If false, the Federation SSO operation might result in a failure with the NoPassive error code, because the IdP will not have been able to identify the user By default set to false RequestedAuthnContext Element indicating how the user should be challenged at the IdP If the SP requests a Federation Authentication Method unknown to the IdP or for which the IdP is not configured, then the Federation SSO flow will result in a failure with the NoAuthnContext error code By default missing NameIDPolicy Element indicating which NameID format the IdP should include in the SAML Assertion If the SP requests a NameID format unknown to the IdP or for which the IdP is not configured, then the Federation SSO flow will result in a failure with the InvalidNameIDPolicy error code If missing, the IdP will generally use the default NameID format configured for this SP partner at the IdP By default missing ProtocolBinding Element indicating which SAML binding should be used by the IdP to redirect the user to the SP with the SAML Assertion Set to Artifact or HTTP-POST By default set to HTTP-POST OIF/SP also allows the administrator to configure the server to: Set which binding should be used by OIF/SP to redirect the user to the IdP with the SAML 2.0 AuthnRequest message: Redirect or HTTP-POST By default set to Redirect Set which binding should be used by OIF/SP to redirect the user to the IdP during logout with SAML 2.0 Logout messages: Redirect or HTTP-POST By default set to Redirect SAML 1.1 The SAML 1.1 specifications do not define a message for the SP to send to the IdP when a Federation SSO operation is started. As such, there is no capability to configure OIF/SP on how to affect the start of the Federation SSO flow. OpenID 2.0 OpenID 2.0 defines several extensions that can be used by the SP/RP to affect how the Federation SSO operation will take place: OpenID request: mode: String indicating if the IdP/OP can visually interact with the user checkid_immediate does not allow the IdP/OP to interact with the user checkid_setup allows user interaction By default set to checkid_setup PAPE Extension: max_auth_age : Integer indicating in seconds the maximum amount of time since when the user authenticated at the IdP. If MaxAuthnAge is bigger that the time since when the user last authenticated at the IdP, then the user must be re-challenged. OIF/SP will set this attribute to 0 if the administrator configured ForceAuthn to true, otherwise this attribute won't be set Default missing preferred_auth_policies Contains a Federation Authentication Method Element indicating how the user should be challenged at the IdP By default missing Only specified in the OpenID request if the IdP/OP supports PAPE in XRDS, if OpenID discovery is used. UI Extension Popup mode Boolean indicating the popup mode is enabled for the Federation SSO By default missing Language Preference String containing the preferred language, set based on the browser's language preferences. By default missing Icon: Boolean indicating if the icon feature is enabled. In that case, the IdP/OP would look at the SP/RP XRDS to determine how to retrieve the icon By default missing Only specified in the OpenID request if the IdP/OP supports UI Extenstion in XRDS, if OpenID discovery is used. ForceAuthn and IsPassive WLST Command OIF/SP provides the WLST configureIdPAuthnRequest() command to set: ForceAuthn as a boolean: In a SAML 2.0 AuthnRequest, the ForceAuthn field will be set to true or false In an OpenID 2.0 request, if ForceAuthn in the configuration was set to true, then the max_auth_age field of the PAPE request will be set to 0, otherwise, max_auth_age won't be set IsPassive as a boolean: In a SAML 2.0 AuthnRequest, the IsPassive field will be set to true or false In an OpenID 2.0 request, if IsPassive in the configuration was set to true, then the mode field of the OpenID request will be set to checkid_immediate, otherwise set to checkid_setup Test In this test, OIF/SP is integrated with a remote SAML 2.0 IdP Partner, with the OOTB configuration. Based on this setup, when OIF/SP starts a Federation SSO flow, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer>   <samlp:NameIDPolicy AllowCreate="true"/></samlp:AuthnRequest> Let's configure OIF/SP for that IdP Partner, so that the SP will require the IdP to re-challenge the user, even if the user is already authenticated: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the configureIdPAuthnRequest() command:configureIdPAuthnRequest(partner="AcmeIdP", forceAuthn="true") Exit the WLST environment:exit() After the changes, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ForceAuthn="true" ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer>   <samlp:NameIDPolicy AllowCreate="true"/></samlp:AuthnRequest> To display or delete the ForceAuthn/IsPassive settings, perform the following operatons: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the configureIdPAuthnRequest() command: To display the ForceAuthn/IsPassive settings on the partnerconfigureIdPAuthnRequest(partner="AcmeIdP", displayOnly="true") To delete the ForceAuthn/IsPassive settings from the partnerconfigureIdPAuthnRequest(partner="AcmeIdP", delete="true") Exit the WLST environment:exit() Requested Fed Authn Method In my earlier "Fed Authentication Method Requests in OIF / SP" article, I discussed how OIF/SP could be configured to request a specific Federation Authentication Method from the IdP when starting a Federation SSO operation, by setting elements in the SSO request message. WLST Command The OIF WLST commands that can be used are: setIdPPartnerProfileRequestAuthnMethod() which will configure the requested Federation Authentication Method in a specific IdP Partner Profile, and accepts the following parameters: partnerProfile: name of the IdP Partner Profile authnMethod: the Federation Authentication Method to request displayOnly: an optional parameter indicating if the method should display the current requested Federation Authentication Method instead of setting it delete: an optional parameter indicating if the method should delete the current requested Federation Authentication Method instead of setting it setIdPPartnerRequestAuthnMethod() which will configure the specified IdP Partner entry with the requested Federation Authentication Method, and accepts the following parameters: partner: name of the IdP Partner authnMethod: the Federation Authentication Method to request displayOnly: an optional parameter indicating if the method should display the current requested Federation Authentication Method instead of setting it delete: an optional parameter indicating if the method should delete the current requested Federation Authentication Method instead of setting it This applies to SAML 2.0 and OpenID 2.0 protocols. See the "Fed Authentication Method Requests in OIF / SP" article for more information. Test In this test, OIF/SP is integrated with a remote SAML 2.0 IdP Partner, with the OOTB configuration. Based on this setup, when OIF/SP starts a Federation SSO flow, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer>   <samlp:NameIDPolicy AllowCreate="true"/></samlp:AuthnRequest> Let's configure OIF/SP for that IdP Partner, so that the SP will request the IdP to use a mechanism mapped to the urn:oasis:names:tc:SAML:2.0:ac:classes:X509 Federation Authentication Method to authenticate the user: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setIdPPartnerRequestAuthnMethod() command:setIdPPartnerRequestAuthnMethod("AcmeIdP", "urn:oasis:names:tc:SAML:2.0:ac:classes:X509") Exit the WLST environment:exit() After the changes, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer>   <samlp:NameIDPolicy AllowCreate="true"/>   <samlp:RequestedAuthnContext Comparison="minimum">      <saml:AuthnContextClassRef xmlns:saml="urn:oasis:names:tc:SAML:2.0:assertion">         urn:oasis:names:tc:SAML:2.0:ac:classes:X509      </saml:AuthnContextClassRef>   </samlp:RequestedAuthnContext></samlp:AuthnRequest> NameID Format The SAML 2.0 protocol allows for the SP to request from the IdP a specific NameID format to be used when the Assertion is issued by the IdP. Note: SAML 1.1 and OpenID 2.0 do not provide such a mechanism Configuring OIF The administrator can configure OIF/SP to request a NameID format in the SAML 2.0 AuthnRequest via: The OAM Administration Console, in the IdP Partner entry The OIF WLST setIdPPartnerNameIDFormat() command that will modify the IdP Partner configuration OAM Administration Console To configure the requested NameID format via the OAM Administration Console, perform the following steps: Go to the OAM Administration Console: http(s)://oam-admin-host:oam-admin-port/oamconsole Navigate to Identity Federation -> Service Provider Administration Open the IdP Partner you wish to modify In the Authentication Request NameID Format dropdown box with one of the values None The NameID format will be set Default Email Address The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:emailAddress X.509 Subject The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:X509SubjectName Windows Name Qualifier The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:WindowsDomainQualifiedName Kerberos The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:kerberos Transient The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:transient Unspecified The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:unspecified Custom In this case, a field would appear allowing the administrator to indicate the custom NameID format to use The NameID format will be set to the specified format Persistent The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:persistent I selected Email Address in this example Save WLST Command To configure the requested NameID format via the OIF WLST setIdPPartnerNameIDFormat() command, perform the following steps: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setIdPPartnerNameIDFormat() command:setIdPPartnerNameIDFormat("PARTNER", "FORMAT", customFormat="CUSTOM") Replace PARTNER with the IdP Partner name Replace FORMAT with one of the following: orafed-none The NameID format will be set Default orafed-emailaddress The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:emailAddress orafed-x509 The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:X509SubjectName orafed-windowsnamequalifier The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:WindowsDomainQualifiedName orafed-kerberos The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:kerberos orafed-transient The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:transient orafed-unspecified The NameID format will be set urn:oasis:names:tc:SAML:1.1:nameid-format:unspecified orafed-custom In this case, a field would appear allowing the administrator to indicate the custom NameID format to use The NameID format will be set to the specified format orafed-persistent The NameID format will be set urn:oasis:names:tc:SAML:2.0:nameid-format:persistent customFormat will need to be set if the FORMAT is set to orafed-custom An example would be:setIdPPartnerNameIDFormat("AcmeIdP", "orafed-emailaddress") Exit the WLST environment:exit() Test In this test, OIF/SP is integrated with a remote SAML 2.0 IdP Partner, with the OOTB configuration. Based on this setup, when OIF/SP starts a Federation SSO flow, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer> <samlp:NameIDPolicy AllowCreate="true"/></samlp:AuthnRequest> After the changes performed either via the OAM Administration Console or via the OIF WLST setIdPPartnerNameIDFormat() command where Email Address would be requested as the NameID Format, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ForceAuthn="false" IsPassive="false" ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer> <samlp:NameIDPolicy Format="urn:oasis:names:tc:SAML:1.1:nameid-format:emailAddress" AllowCreate="true"/></samlp:AuthnRequest> Protocol Binding The SAML 2.0 specifications define a way for the SP to request which binding should be used by the IdP to redirect the user to the SP with the SAML 2.0 Assertion: the ProtocolBinding attribute indicates the binding the IdP should use. It is set to: Either urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST for HTTP-POST Or urn:oasis:names:tc:SAML:2.0:bindings:Artifact for Artifact The SAML 2.0 specifications also define different ways to redirect the user from the SP to the IdP with the SAML 2.0 AuthnRequest message, as the SP can send the message: Either via HTTP Redirect Or HTTP POST (Other bindings can theoretically be used such as Artifact, but these are not used in practice) Configuring OIF OIF can be configured: Via the OAM Administration Console or the OIF WLST configureSAMLBinding() command to set the Assertion Response binding to be used Via the OIF WLST configureSAMLBinding() command to indicate how the SAML AuthnRequest message should be sent Note: the binding for sending the SAML 2.0 AuthnRequest message will also be used to send the SAML 2.0 LogoutRequest and LogoutResponse messages. OAM Administration Console To configure the SSO Response/Assertion Binding via the OAM Administration Console, perform the following steps: Go to the OAM Administration Console: http(s)://oam-admin-host:oam-admin-port/oamconsole Navigate to Identity Federation -> Service Provider Administration Open the IdP Partner you wish to modify Check the "HTTP POST SSO Response Binding" box to request the IdP to return the SSO Response via HTTP POST, otherwise uncheck it to request artifact Save WLST Command To configure the SSO Response/Assertion Binding as well as the AuthnRequest Binding via the OIF WLST configureSAMLBinding() command, perform the following steps: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the configureSAMLBinding() command:configureSAMLBinding("PARTNER", "PARTNER_TYPE", binding, ssoResponseBinding="httppost") Replace PARTNER with the Partner name Replace PARTNER_TYPE with the Partner type (idp or sp) Replace binding with the binding to be used to send the AuthnRequest and LogoutRequest/LogoutResponse messages (should be httpredirect in most case; default) httppost for HTTP-POST binding httpredirect for HTTP-Redirect binding Specify optionally ssoResponseBinding to indicate how the SSO Assertion should be sent back httppost for HTTP-POST binding artifactfor for Artifact binding An example would be:configureSAMLBinding("AcmeIdP", "idp", "httpredirect", ssoResponseBinding="httppost") Exit the WLST environment:exit() Test In this test, OIF/SP is integrated with a remote SAML 2.0 IdP Partner, with the OOTB configuration which requests HTTP-POST from the IdP to send the SSO Assertion. Based on this setup, when OIF/SP starts a Federation SSO flow, the following SAML 2.0 AuthnRequest would be generated: <samlp:AuthnRequest ProtocolBinding="urn:oasis:names:tc:SAML:2.0:bindings:HTTP-POST" ID="id-E4BOT7lwbYK56lO57dBaqGUFq01WJSjAHiSR60Q4" Version="2.0" IssueInstant="2014-04-01T21:39:14Z" Destination="https://acme.com/saml20/sso">   <saml:Issuer Format="urn:oasis:names:tc:SAML:2.0:nameid-format:entity">https://sp.com/oam/fed</saml:Issuer>   <samlp:NameIDPolicy AllowCreate="true"/></samlp:AuthnRequest> In the next article, I will cover the various crypto configuration properties in OIF that are used to affect the Federation SSO exchanges.Cheers,Damien Carru

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  • Basic Spatial Data with SQL Server and Entity Framework 5.0

    - by Rick Strahl
    In my most recent project we needed to do a bit of geo-spatial referencing. While spatial features have been in SQL Server for a while using those features inside of .NET applications hasn't been as straight forward as could be, because .NET natively doesn't support spatial types. There are workarounds for this with a few custom project like SharpMap or a hack using the Sql Server specific Geo types found in the Microsoft.SqlTypes assembly that ships with SQL server. While these approaches work for manipulating spatial data from .NET code, they didn't work with database access if you're using Entity Framework. Other ORM vendors have been rolling their own versions of spatial integration. In Entity Framework 5.0 running on .NET 4.5 the Microsoft ORM finally adds support for spatial types as well. In this post I'll describe basic geography features that deal with single location and distance calculations which is probably the most common usage scenario. SQL Server Transact-SQL Syntax for Spatial Data Before we look at how things work with Entity framework, lets take a look at how SQL Server allows you to use spatial data to get an understanding of the underlying semantics. The following SQL examples should work with SQL 2008 and forward. Let's start by creating a test table that includes a Geography field and also a pair of Long/Lat fields that demonstrate how you can work with the geography functions even if you don't have geography/geometry fields in the database. Here's the CREATE command:CREATE TABLE [dbo].[Geo]( [id] [int] IDENTITY(1,1) NOT NULL, [Location] [geography] NULL, [Long] [float] NOT NULL, [Lat] [float] NOT NULL ) Now using plain SQL you can insert data into the table using geography::STGeoFromText SQL CLR function:insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.527200 45.712113)', 4326), -121.527200, 45.712113 ) insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.517265 45.714240)', 4326), -121.517265, 45.714240 ) insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.511536 45.714825)', 4326), -121.511536, 45.714825) The STGeomFromText function accepts a string that points to a geometric item (a point here but can also be a line or path or polygon and many others). You also need to provide an SRID (Spatial Reference System Identifier) which is an integer value that determines the rules for how geography/geometry values are calculated and returned. For mapping/distance functionality you typically want to use 4326 as this is the format used by most mapping software and geo-location libraries like Google and Bing. The spatial data in the Location field is stored in binary format which looks something like this: Once the location data is in the database you can query the data and do simple distance computations very easily. For example to calculate the distance of each of the values in the database to another spatial point is very easy to calculate. Distance calculations compare two points in space using a direct line calculation. For our example I'll compare a new point to all the points in the database. Using the Location field the SQL looks like this:-- create a source point DECLARE @s geography SET @s = geography:: STGeomFromText('POINT(-121.527200 45.712113)' , 4326); --- return the ids select ID, Location as Geo , Location .ToString() as Point , @s.STDistance( Location) as distance from Geo order by distance The code defines a new point which is the base point to compare each of the values to. You can also compare values from the database directly, but typically you'll want to match a location to another location and determine the difference for which you can use the geography::STDistance function. This query produces the following output: The STDistance function returns the straight line distance between the passed in point and the point in the database field. The result for SRID 4326 is always in meters. Notice that the first value passed was the same point so the difference is 0. The other two points are two points here in town in Hood River a little ways away - 808 and 1256 meters respectively. Notice also that you can order the result by the resulting distance, which effectively gives you results that are ordered radially out from closer to further away. This is great for searches of points of interest near a central location (YOU typically!). These geolocation functions are also available to you if you don't use the Geography/Geometry types, but plain float values. It's a little more work, as each point has to be created in the query using the string syntax, but the following code doesn't use a geography field but produces the same result as the previous query.--- using float fields select ID, geography::STGeomFromText ('POINT(' + STR (long, 15,7 ) + ' ' + Str(lat ,15, 7) + ')' , 4326), geography::STGeomFromText ('POINT(' + STR (long, 15,7 ) + ' ' + Str(lat ,15, 7) + ')' , 4326). ToString(), @s.STDistance( geography::STGeomFromText ('POINT(' + STR(long ,15, 7) + ' ' + Str(lat ,15, 7) + ')' , 4326)) as distance from geo order by distance Spatial Data in the Entity Framework Prior to Entity Framework 5.0 on .NET 4.5 consuming of the data above required using stored procedures or raw SQL commands to access the spatial data. In Entity Framework 5 however, Microsoft introduced the new DbGeometry and DbGeography types. These immutable location types provide a bunch of functionality for manipulating spatial points using geometry functions which in turn can be used to do common spatial queries like I described in the SQL syntax above. The DbGeography/DbGeometry types are immutable, meaning that you can't write to them once they've been created. They are a bit odd in that you need to use factory methods in order to instantiate them - they have no constructor() and you can't assign to properties like Latitude and Longitude. Creating a Model with Spatial Data Let's start by creating a simple Entity Framework model that includes a Location property of type DbGeography: public class GeoLocationContext : DbContext { public DbSet<GeoLocation> Locations { get; set; } } public class GeoLocation { public int Id { get; set; } public DbGeography Location { get; set; } public string Address { get; set; } } That's all there's to it. When you run this now against SQL Server, you get a Geography field for the Location property, which looks the same as the Location field in the SQL examples earlier. Adding Spatial Data to the Database Next let's add some data to the table that includes some latitude and longitude data. An easy way to find lat/long locations is to use Google Maps to pinpoint your location, then right click and click on What's Here. Click on the green marker to get the GPS coordinates. To add the actual geolocation data create an instance of the GeoLocation type and use the DbGeography.PointFromText() factory method to create a new point to assign to the Location property:[TestMethod] public void AddLocationsToDataBase() { var context = new GeoLocationContext(); // remove all context.Locations.ToList().ForEach( loc => context.Locations.Remove(loc)); context.SaveChanges(); var location = new GeoLocation() { // Create a point using native DbGeography Factory method Location = DbGeography.PointFromText( string.Format("POINT({0} {1})", -121.527200,45.712113) ,4326), Address = "301 15th Street, Hood River" }; context.Locations.Add(location); location = new GeoLocation() { Location = CreatePoint(45.714240, -121.517265), Address = "The Hatchery, Bingen" }; context.Locations.Add(location); location = new GeoLocation() { // Create a point using a helper function (lat/long) Location = CreatePoint(45.708457, -121.514432), Address = "Kaze Sushi, Hood River" }; context.Locations.Add(location); location = new GeoLocation() { Location = CreatePoint(45.722780, -120.209227), Address = "Arlington, OR" }; context.Locations.Add(location); context.SaveChanges(); } As promised, a DbGeography object has to be created with one of the static factory methods provided on the type as the Location.Longitude and Location.Latitude properties are read only. Here I'm using PointFromText() which uses a "Well Known Text" format to specify spatial data. In the first example I'm specifying to create a Point from a longitude and latitude value, using an SRID of 4326 (just like earlier in the SQL examples). You'll probably want to create a helper method to make the creation of Points easier to avoid that string format and instead just pass in a couple of double values. Here's my helper called CreatePoint that's used for all but the first point creation in the sample above:public static DbGeography CreatePoint(double latitude, double longitude) { var text = string.Format(CultureInfo.InvariantCulture.NumberFormat, "POINT({0} {1})", longitude, latitude); // 4326 is most common coordinate system used by GPS/Maps return DbGeography.PointFromText(text, 4326); } Using the helper the syntax becomes a bit cleaner, requiring only a latitude and longitude respectively. Note that my method intentionally swaps the parameters around because Latitude and Longitude is the common format I've seen with mapping libraries (especially Google Mapping/Geolocation APIs with their LatLng type). When the context is changed the data is written into the database using the SQL Geography type which looks the same as in the earlier SQL examples shown. Querying Once you have some location data in the database it's now super easy to query the data and find out the distance between locations. A common query is to ask for a number of locations that are near a fixed point - typically your current location and order it by distance. Using LINQ to Entities a query like this is easy to construct:[TestMethod] public void QueryLocationsTest() { var sourcePoint = CreatePoint(45.712113, -121.527200); var context = new GeoLocationContext(); // find any locations within 5 kilometers ordered by distance var matches = context.Locations .Where(loc => loc.Location.Distance(sourcePoint) < 5000) .OrderBy( loc=> loc.Location.Distance(sourcePoint) ) .Select( loc=> new { Address = loc.Address, Distance = loc.Location.Distance(sourcePoint) }); Assert.IsTrue(matches.Count() > 0); foreach (var location in matches) { Console.WriteLine("{0} ({1:n0} meters)", location.Address, location.Distance); } } This example produces: 301 15th Street, Hood River (0 meters)The Hatchery, Bingen (809 meters)Kaze Sushi, Hood River (1,074 meters)   The first point in the database is the same as my source point I'm comparing against so the distance is 0. The other two are within the 5 mile radius, while the Arlington location which is 65 miles or so out is not returned. The result is ordered by distance from closest to furthest away. In the code, I first create a source point that is the basis for comparison. The LINQ query then selects all locations that are within 5km of the source point using the Location.Distance() function, which takes a source point as a parameter. You can either use a pre-defined value as I'm doing here, or compare against another database DbGeography property (say when you have to points in the same database for things like routes). What's nice about this query syntax is that it's very clean and easy to read and understand. You can calculate the distance and also easily order by the distance to provide a result that shows locations from closest to furthest away which is a common scenario for any application that places a user in the context of several locations. It's now super easy to accomplish this. Meters vs. Miles As with the SQL Server functions, the Distance() method returns data in meters, so if you need to work with miles or feet you need to do some conversion. Here are a couple of helpers that might be useful (can be found in GeoUtils.cs of the sample project):/// <summary> /// Convert meters to miles /// </summary> /// <param name="meters"></param> /// <returns></returns> public static double MetersToMiles(double? meters) { if (meters == null) return 0F; return meters.Value * 0.000621371192; } /// <summary> /// Convert miles to meters /// </summary> /// <param name="miles"></param> /// <returns></returns> public static double MilesToMeters(double? miles) { if (miles == null) return 0; return miles.Value * 1609.344; } Using these two helpers you can query on miles like this:[TestMethod] public void QueryLocationsMilesTest() { var sourcePoint = CreatePoint(45.712113, -121.527200); var context = new GeoLocationContext(); // find any locations within 5 miles ordered by distance var fiveMiles = GeoUtils.MilesToMeters(5); var matches = context.Locations .Where(loc => loc.Location.Distance(sourcePoint) <= fiveMiles) .OrderBy(loc => loc.Location.Distance(sourcePoint)) .Select(loc => new { Address = loc.Address, Distance = loc.Location.Distance(sourcePoint) }); Assert.IsTrue(matches.Count() > 0); foreach (var location in matches) { Console.WriteLine("{0} ({1:n1} miles)", location.Address, GeoUtils.MetersToMiles(location.Distance)); } } which produces: 301 15th Street, Hood River (0.0 miles)The Hatchery, Bingen (0.5 miles)Kaze Sushi, Hood River (0.7 miles) Nice 'n simple. .NET 4.5 Only Note that DbGeography and DbGeometry are exclusive to Entity Framework 5.0 (not 4.4 which ships in the same NuGet package or installer) and requires .NET 4.5. That's because the new DbGeometry and DbGeography (and related) types are defined in the 4.5 version of System.Data.Entity which is a CLR assembly and is only updated by major versions of .NET. Why this decision was made to add these types to System.Data.Entity rather than to the frequently updated EntityFramework assembly that would have possibly made this work in .NET 4.0 is beyond me, especially given that there are no native .NET framework spatial types to begin with. I find it also odd that there is no native CLR spatial type. The DbGeography and DbGeometry types are specific to Entity Framework and live on those assemblies. They will also work for general purpose, non-database spatial data manipulation, but then you are forced into having a dependency on System.Data.Entity, which seems a bit silly. There's also a System.Spatial assembly that's apparently part of WCF Data Services which in turn don't work with Entity framework. Another example of multiple teams at Microsoft not communicating and implementing the same functionality (differently) in several different places. Perplexed as a I may be, for EF specific code the Entity framework specific types are easy to use and work well. Working with pre-.NET 4.5 Entity Framework and Spatial Data If you can't go to .NET 4.5 just yet you can also still use spatial features in Entity Framework, but it's a lot more work as you can't use the DbContext directly to manipulate the location data. You can still run raw SQL statements to write data into the database and retrieve results using the same TSQL syntax I showed earlier using Context.Database.ExecuteSqlCommand(). Here's code that you can use to add location data into the database:[TestMethod] public void RawSqlEfAddTest() { string sqlFormat = @"insert into GeoLocations( Location, Address) values ( geography::STGeomFromText('POINT({0} {1})', 4326),@p0 )"; var sql = string.Format(sqlFormat,-121.527200, 45.712113); Console.WriteLine(sql); var context = new GeoLocationContext(); Assert.IsTrue(context.Database.ExecuteSqlCommand(sql,"301 N. 15th Street") > 0); } Here I'm using the STGeomFromText() function to add the location data. Note that I'm using string.Format here, which usually would be a bad practice but is required here. I was unable to use ExecuteSqlCommand() and its named parameter syntax as the longitude and latitude parameters are embedded into a string. Rest assured it's required as the following does not work:string sqlFormat = @"insert into GeoLocations( Location, Address) values ( geography::STGeomFromText('POINT(@p0 @p1)', 4326),@p2 )";context.Database.ExecuteSqlCommand(sql, -121.527200, 45.712113, "301 N. 15th Street") Explicitly assigning the point value with string.format works however. There are a number of ways to query location data. You can't get the location data directly, but you can retrieve the point string (which can then be parsed to get Latitude and Longitude) and you can return calculated values like distance. Here's an example of how to retrieve some geo data into a resultset using EF's and SqlQuery method:[TestMethod] public void RawSqlEfQueryTest() { var sqlFormat = @" DECLARE @s geography SET @s = geography:: STGeomFromText('POINT({0} {1})' , 4326); SELECT Address, Location.ToString() as GeoString, @s.STDistance( Location) as Distance FROM GeoLocations ORDER BY Distance"; var sql = string.Format(sqlFormat, -121.527200, 45.712113); var context = new GeoLocationContext(); var locations = context.Database.SqlQuery<ResultData>(sql); Assert.IsTrue(locations.Count() > 0); foreach (var location in locations) { Console.WriteLine(location.Address + " " + location.GeoString + " " + location.Distance); } } public class ResultData { public string GeoString { get; set; } public double Distance { get; set; } public string Address { get; set; } } Hopefully you don't have to resort to this approach as it's fairly limited. Using the new DbGeography/DbGeometry types makes this sort of thing so much easier. When I had to use code like this before I typically ended up retrieving data pks only and then running another query with just the PKs to retrieve the actual underlying DbContext entities. This was very inefficient and tedious but it did work. Summary For the current project I'm working on we actually made the switch to .NET 4.5 purely for the spatial features in EF 5.0. This app heavily relies on spatial queries and it was worth taking a chance with pre-release code to get this ease of integration as opposed to manually falling back to stored procedures or raw SQL string queries to return spatial specific queries. Using native Entity Framework code makes life a lot easier than the alternatives. It might be a late addition to Entity Framework, but it sure makes location calculations and storage easy. Where do you want to go today? ;-) Resources Download Sample Project© Rick Strahl, West Wind Technologies, 2005-2012Posted in ADO.NET  Sql Server  .NET   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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