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  • ASP.NET MVC 4/Web API Single Page App for Mobile Devices ... Needs Authentication

    - by lmttag
    We have developed an ASP.NET MVC 4/Web API single page, mobile website (also using jQuery Mobile) that is intended to be accessed only from mobile devices (e.g., iPads, iPhones, Android tables and phones, etc.), not desktop browsers. This mobile website will be hosted internally, like an intranet site. However, since we’re accessing it from mobile devices, we can’t use Windows authentication. We still need to know which user (and their role) is logging in to the mobile website app. We tried simply using ASP.NET’s forms authentication and membership provider, but couldn’t get it working exactly the way we wanted. What we need is for the user to be prompted for a user name and password only on the first time they access the site on their mobile device. After they enter a correct user name and password and have been authenticated once, each subsequent time they access the site they should just go right in. They shouldn’t have to re-enter their credentials (i.e., something needs to be saved locally to each device to identify the user after the first time). This is where we had troubles. Everything worked as expected the first time. That is, the user was prompted to enter a user name and password, and, after doing that, was authenticated and allowed into the site. The problem is every time after the browser was closed on the mobile device, the device and user were not know and the user had to re-enter user name and password. We tried lots of things too. We tried setting persistent cookies in JavaScript. No good. The cookies weren’t there to be read the second time. We tried manually setting persistent cookies from ASP.NET. No good. We, of course, used FormsAuthentication.SetAuthCookie(model.UserName, true); as part of the form authentication framework. No good. We tried using HTML5 local storage. No good. No matter what we tried, if the user was on a mobile device, they would have to log in every single time. (Note: we’ve tried on an iPad and iPhone running both iOS 5.1 and 6.0, with Safari configure to allow cookies, and we’ve tried on Android 2.3.4.) Is there some trick to getting a scenario like this working? Or, do we have to write some sort of custom authentication mechanism? If so, how? And, what? Or, should we use something like claims-based authentication and WIF? Or??? Any help is appreciated. Thanks!

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  • Akka framework support for finding duplicate messages

    - by scala_is_awesome
    I'm trying to build a high-performance distributed system with Akka and Scala. If a message requesting an expensive (and side-effect-free) computation arrives, and the exact same computation has already been requested before, I want to avoid computing the result again. If the computation requested previously has already completed and the result is available, I can cache it and re-use it. However, the time window in which duplicate computation can be requested may be arbitrarily small. e.g. I could get a thousand or a million messages requesting the same expensive computation at the same instant for all practical purposes. There is a commercial product called Gigaspaces that supposedly handles this situation. However there seems to be no framework support for dealing with duplicate work requests in Akka at the moment. Given that the Akka framework already has access to all the messages being routed through the framework, it seems that a framework solution could make a lot of sense here. Here is what I am proposing for the Akka framework to do: 1. Create a trait to indicate a type of messages (say, "ExpensiveComputation" or something similar) that are to be subject to the following caching approach. 2. Smartly (hashing etc.) identify identical messages received by (the same or different) actors within a user-configurable time window. Other options: select a maximum buffer size of memory to be used for this purpose, subject to (say LRU) replacement etc. Akka can also choose to cache only the results of messages that were expensive to process; the messages that took very little time to process can be re-processed again if needed; no need to waste precious buffer space caching them and their results. 3. When identical messages (received within that time window, possibly "at the same time instant") are identified, avoid unnecessary duplicate computations. The framework would do this automatically, and essentially, the duplicate messages would never get received by a new actor for processing; they would silently vanish and the result from processing it once (whether that computation was already done in the past, or ongoing right then) would get sent to all appropriate recipients (immediately if already available, and upon completion of the computation if not). Note that messages should be considered identical even if the "reply" fields are different, as long as the semantics/computations they represent are identical in every other respect. Also note that the computation should be purely functional, i.e. free from side-effects, for the caching optimization suggested to work and not change the program semantics at all. If what I am suggesting is not compatible with the Akka way of doing things, and/or if you see some strong reasons why this is a very bad idea, please let me know. Thanks, Is Awesome, Scala

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  • RaphaelJS HTML5 Library pathIntersection() bug or alternative optimisation (screenshots)

    - by user1236048
    I have a chart generated using RaphaelJS library. It is just on long path: M 50 122 L 63.230769230769226 130 L 76.46153846153845 130 L 89.6923076923077 128 L 102.92307692307692 56 L 116.15384615384615 106 L 129.3846153846154 88 L 142.6153846153846 114 L 155.84615384615384 52 L 169.07692307692307 30 L 182.3076923076923 62 L 195.53846153846152 130 L 208.76923076923077 74 L 222 130 L 235.23076923076923 66 L 248.46153846153845 102 L 261.6923076923077 32 L 274.9230769230769 130 L 288.15384615384613 130 L 301.38461538461536 32 L 314.6153846153846 86 L 327.8461538461538 130 L 341.07692307692304 70 L 354.30769230769226 130 L 367.53846153846155 102 L 380.7692307692308 120 L 394 112 L 407.2307692307692 68 L 420.46153846153845 48 L 433.6923076923077 92 L 446.9230769230769 128 L 460.15384615384613 110 L 473.38461538461536 78 L 486.6153846153846 130 L 499.8461538461538 56 L 513.0769230769231 116 L 526.3076923076923 80 L 539.5384615384614 58 L 552.7692307692307 40 L 566 130 L 579.2307692307692 94 L 592.4615384615385 64 L 605.6923076923076 122 L 618.9230769230769 98 L 632.1538461538461 120 L 645.3846153846154 70 L 658.6153846153845 82 L 671.8461538461538 76 L 685.0769230769231 124 L 698.3076923076923 110 L 711.5384615384615 94 L 724.7692307692307 130 L 738 130 L 751.2307692307692 66 L 764.4615384615385 118 L 777.6923076923076 70 L 790.9230769230769 130 L 804.1538461538461 44 L 817.3846153846154 130 L 830.6153846153845 36 L 843.8461538461538 92 L 857.076923076923 130 L 870.3076923076923 76 L 883.5384615384614 130 L 896.7692307692307 60 L 910 88 Also below these chart I have a jqueryUI slider of the same width (860px) and centered with the chart. I want when I move the slider to move a dot on the chart accordingly with the slider position. See attached screenshot: As you can see it seems to work fine. I've implemented this behaviour using the pathIntersection() method. On the slide event at each ui.value (x coordinate) I intersect my chartPath (the one from above) with a vertical straight line at the x coordinate. But still there are some problems. One of them is that it runs very hard, and it kinda freezes sometimes.. and very weird sometimes it doesn't seem to intersect at all even it should.. I'll example below 2 cases I identified: M 499.8461538461538 0 L 499.8461538461538 140 M 910 0 L 910 140 Could you please explain why this intersect behaviour happens (it should return a dot).. and the worst part it seems like it happens randomly.. if I use another chartdata. Also if you can identify another (better) solution to syncronise the slider position with the dot on the chart.. would be perfect. I thought about using Element.getPointAtLength(length), but I don't know how. I think I should save the pathSegments and for each to compute the start Length and the finish Length.

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  • Technical non-terminating condition in a loop

    - by Snarfblam
    Most of us know that a loop should not have a non-terminating condition. For example, this C# loop has a non-terminating condition: any even value of i. This is an obvious logic error. void CountByTwosStartingAt(byte i) { // If i is even, it never exceeds 254 for(; i < 255; i += 2) { Console.WriteLine(i); } } Sometimes there are edge cases that are extremely unlikeley, but technically constitute non-exiting conditions (stack overflows and out-of-memory errors aside). Suppose you have a function that counts the number of sequential zeros in a stream: int CountZeros(Stream s) { int total = 0; while(s.ReadByte() == 0) total++; return total; } Now, suppose you feed it this thing: class InfiniteEmptyStream:Stream { // ... Other members ... public override int Read(byte[] buffer, int offset, int count) { Array.Clear(buffer, offset, count); // Output zeros return count; // Never returns -1 (end of stream) } } Or more realistically, maybe a stream that returns data from external hardware, which in certain cases might return lots of zeros (such as a game controller sitting on your desk). Either way we have an infinite loop. This particular non-terminating condition stands out, but sometimes they don't. A completely real-world example as in an app I'm writing. An endless stream of zeros will be deserialized into infinite "empty" objects (until the collection class or GC throws an exception because I've exceeded two billion items). But this would be a completely unexpected circumstance (considering my data source). How important is it to have absolutely no non-terminating conditions? How much does this affect "robustness?" Does it matter if they are only "theoretically" non-terminating (is it okay if an exception represents an implicit terminating condition)? Does it matter whether the app is commercial? If it is publicly distributed? Does it matter if the problematic code is in no way accessible through a public interface/API? Edit: One of the primary concerns I have is unforseen logic errors that can create the non-terminating condition. If, as a rule, you ensure there are no non-terminating conditions, you can identify or handle these logic errors more gracefully, but is it worth it? And when? This is a concern orthogonal to trust.

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  • Send multidimensional array to php with jQuery

    - by robertdd
    hy, i have a little, big problem here :) after i upload some images i get a list with all the images. I have some jQuery function for rotate, duplicate, delete, shuffle images! when i select a image and hit delete i send a post to php with the alt="" value of the image,i identify the picture and edit. I want to make a save button, Instead of sending a post every time i rotate a image, better send a post after editing the list of images with an array that contains all data? my php array after upload looks like this: [files] => Array ( [lcxkijgr] => lcxkijgr.jpg [xcewxpfv] => xcewxpfv.jpg [rtiurwxf] => rtiurwxf.jpg [gsbxdsdc] => gsbxdsdc.jpg ) say that I uploaded 4 pictures, firs picture i rotate 90 degrees second i want to duplicate third i rotate 270 degrees and the fourth image i delete i can do all this only with jQuery, but on the server the images are the same, after a refresh the images are the same this is the list with the images: <div class="upimage"> <ul id="upimagesQueue"> <li id="upimagesHPVEJM"> <a href="javascript:jQuery('#upimagesHPVEJM').showlargeimage('HPVEJM')"> <img alt="lcxkijgr" src="uploads/s6id9r9icnp8q9102h8md9kfd7/lcxkijgr.jpg?1272087830477" id="HPVEJM" style="display: block;" > </a> </li> <li id="upimagesSTCSAV"> <a href="javascript:jQuery('#upimagesSTCSAV').showlargeimage('STCSAV')"> <img alt="xcewxpfv" src="uploads/s6id9r9icnp8q9102h8md9kfd7/xcewxpfv.jpg?1272087831360" id="STCSAV" style="display: block;" > </a> </li> <li id="upimagesBFPUEQ"> <a href="javascript:jQuery('#upimagesBFPUEQ').showlargeimage('BFPUEQ')"> <img alt="rtiurwxf" src="uploads/s6id9r9icnp8q9102h8md9kfd7/rtiurwxf.jpg?1272087832162" id="BFPUEQ" style="display: block;" > </a> </li> <li id="upimagesRKXNSV"> <a href="javascript:jQuery('#upimagesRKXNSV').showlargeimage('RKXNSV')"> <img alt="gsbxdsdc" src="uploads/s6id9r9icnp8q9102h8md9kfd7/gsbxdsdc.jpg?1272087832957" id="RKXNSV" style="display: block;"> </a> </li> <ul> </div> is ok if i make one array like this: array{ imgFromLi = array(img1,img2,img3,img4,img5,img6) rotate = array{img1=90, img2=270, img3=90} delete = array{img4,img5,img6} duplicate = array{img2, img3} } how i can make/send/cache this array?? sorry for my very bad english

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  • issues regarding UAC prompt

    - by peter
    I want to implement a UAC prompt for an application in visualc++ the operating system is 32bit x7460(2processor) Windowsserver 2008 the exe is myproject.exe through manifest.. Here for testing i wl build the application in Windows XP OS and copy the exe in to system containg the Windowsserver 2008 machine and replace it So what i did is i added a manifest like this name of that is myproject.exe.manifest My project has 3 folders like Headerfile,Resourcefile and Source file.I added this manifest(myproject.exe.manifest) in the Sourcefile folder containing other cpp and c code <?xml version="1.0" encoding="UTF-8" standalone="yes"?> <assembly xmlns="urn:schemas-microsoft-com:asm.v1" manifestVersion="1.0"> <assemblyIdentity version="4.0" processorArchitecture="X7460" name="myproject" type="win32"/> <description>myproject Problem</description> <!-- Identify the application security requirements. --> <trustInfo xmlns="urn:schemas-microsoft-com:asm.v2"> <security> <requestedPrivileges> <requestedExecutionLevel level="requireAdministrator" uiAccess="false"/> </requestedPrivileges> </security> </trustInfo> </assembly> then i added this line of code in Resourcefile(.rc).Means one header file is there(Myproject.h).I added the line of code there #define MANIFEST_RESOURCE_ID 1 MANIFEST_RESOURCE_ID RT_MANIFEST "myproject.exe.manifest" Finally i did the following step Under Project, select Properties. 3. In Properties, select Manifest Tool, and then select Input and Output. 4. Add in the name of your application manifest file under Additional manifest files. 5. Rebuild your application. But i am getting lot of Syntax errors Is there any problems in the way which i followed.If i commented the line #define MANIFEST_RESOURCE_ID 1 MANIFEST_RESOURCE_ID RT_MANIFEST "myproject.exe.manifest" which added in Myproject.h for adding values in .rc file there willnot any error other than this general error c1010070: Failed to load and parse the manifest. The system cannot find the file specified. .\myproject.exe.manifest How to enable UAC prompt through programming

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  • Different behavior of reflected generic delegates with and without debugger

    - by Andrew_B
    Hello. We have encountered some strange things while calling reflected generic delegates. In some cases with attatched debuger we can make impossible call, while without debugger we cannot catch any exception and application fastfails. Here is the code: using System; using System.Windows.Forms; using System.Reflection; namespace GenericDelegate { public partial class Form1 : Form { public Form1() { InitializeComponent(); } private delegate Class2 Delegate1(); private void button1_Click(object sender, EventArgs e) { MethodInfo mi = typeof (Class1<>).GetMethod("GetClass", BindingFlags.NonPublic | BindingFlags.Static); if (mi != null) { Delegate1 del = (Delegate1) Delegate.CreateDelegate(typeof (Delegate1), mi); MessageBox.Show("1"); try { del(); } catch (Exception) { MessageBox.Show("No, I can`t catch it"); } MessageBox.Show("2"); mi.Invoke(null, new object[] {});//It's Ok, we'll get exception here MessageBox.Show("3"); } } class Class2 { } class Class1<T> : Class2 { internal static Class2 GetClass() { Type type = typeof(T); MessageBox.Show("Type name " + type.FullName +" Type: " + type + " Assembly " + type.Assembly); return new Class1<T>(); } } } } There are two problems: Behavior differs with debugger and without You cannot catch this error without debugger by clr tricks. It's just not the clr exception. There are memory acces vialation, reading zero pointer inside of internal code. Use case: You develop something like plugins system for your app. You read external assembly, find suitable method in some type, and execute it. And we just forgot about that we need to check up is the type generic or not. Under VS (and .net from 2.0 to 4.0) everything works fine. Called function does not uses static context of generic type and type parameters. But without VS application fails with no sound. We even cannot identify call stack attaching debuger. Tested with .net 4.0 The question is why VS catches but runtime do not?

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  • C++ problem with string stream istringstream

    - by user69514
    I am reading a file in the following format 1001 16000 300 12.50 2002 24000 360 10.50 3003 30000 300 9.50 where the items are: loan id, principal, months, interest rate. I'm not sure what it is that I am doing wrong with my input string stream, but I am not reading the values correctly because only the loan id is read correctly. Everything else is zero. Sorry this is a homework, but I just wanted to know if you could help me identify my error. if( inputstream.is_open() ){ /** print the results **/ cout << fixed << showpoint << setprecision(2); cout << "ID " << "\tPrincipal" << "\tDuration" << "\tInterest" << "\tPayment" <<"\tTotal Payment" << endl; cout << "---------------------------------------------------------------------------------------------" << endl; /** assign line read while we haven't reached end of file **/ string line; istringstream instream; while( inputstream >> line ){ instream.clear(); instream.str(line); /** assing values **/ instream >> loanid >> principal >> duration >> interest; /** compute monthly payment **/ double ratem = interest / 1200.0; double expm = (1.0 + ratem); payment = (ratem * pow(expm, duration) * principal) / (pow(expm, duration) - 1.0); /** computer total payment **/ totalPayment = payment * duration; /** print out calculations **/ cout << loanid << "\t$" << principal <<"\t" << duration << "mo" << "\t" << interest << "\t$" << payment << "\t$" << totalPayment << endl; } }

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  • How to eliminate one of my extra DropDownLists in ASP.NET?

    - by salvationishere
    I'm developing a C#/SQL web app in VS 2008 but for some reason I have one extra DropDownList. The very first dropdownlist displaying is empty. Can you help me identify the cause of this behavior? I'm baffled! An excerpt of my code is below. private DropDownList[] newcol; // Add DropDownList Control to Placeholder private DropDownList[] CreateDropDownLists() { DropDownList[] dropDowns = new DropDownList[NumberOfControls]; for (int counter = 0; counter < NumberOfControls; counter++) { DropDownList ddl = new DropDownList(); SqlDataReader dr2 = ADONET_methods.DisplayTableColumns(targettable); ddl.ID = "DropDownListID" + counter.ToString(); int NumControls = targettable.Length; DataTable dt = new DataTable(); dt.Load(dr2); ddl.DataValueField = "COLUMN_NAME"; ddl.DataTextField = "COLUMN_NAME"; ddl.DataSource = dt; ddl.SelectedIndexChanged += new EventHandler(ddlList_SelectedIndexChanged); ddl.DataBind(); ddl.AutoPostBack = true; ddl.EnableViewState = true; //Preserves View State info on Postbacks //ddlList.Style["position"] = "absolute"; //ddl.Style["top"] = 80 + "px"; //ddl.Style["left"] = 0 + "px"; dr2.Close(); dropDowns[counter] = ddl; } return dropDowns; } protected void ddlList_SelectedIndexChanged(object sender, EventArgs e) { DropDownList ddl = (DropDownList)sender; string ID = ddl.ID; } //Create display panel private void CreateDisplayPanel() { btnSubmit.Style.Add("top", "auto"); btnSubmit.Style.Add("left", "auto"); btnSubmit.Style.Add("position", "absolute"); newcol = CreateDropDownLists(); for (int counter = 0; counter < NumberOfControls; counter++) { pnlDisplayData.Controls.Add(newcol[counter]); pnlDisplayData.Controls.Add(new LiteralControl("<br><br><br>")); pnlDisplayData.Visible = true; pnlDisplayData.FindControl(newcol[counter].ID); } }

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  • how to invoke an activity of a library project from an android apps

    - by Austin
    I have an open source android code that I need to use in my android apps. It has all the source code as well as resource files, manifest files and class path. It can be compiled as a separate android apps. I have constraints for using the open source. 1. I can't change a single line of code. 2. I can't use it as a separate apps. These constraints are non negotiable. What I have done is I have compiled the open source as class library(in Eclipse: Project Properties-Android- Tick check box Is Library). This has resulted in generation of .class files(in bin) for the java files and resource files. This open source has an android activity that i want to open from my application. So I have linked the directory of these sets of class files in the source section of my java build path( in .classpath). I have declared the activity in my manifest file with proper action intent filters. Now when I am trying to call activity from my code, its not working. Cleaning and rebuilding doesn't help. However, if I build the open source project and my apps in the same workspace of eclipse and link the open source in my apps in exact same manner it works fine. I am not able to identify the difference. All settings seems to be same(all files are identical in both the cases). But only in the second case it works. I have tried it as jar file also. I have build the open source as project library and exported it into a jar file(excluding manifest file). But in that case I am getting the following error UNEXPECTED TOP-LEVEL EXCEPTION: java.lang.IllegalArgumentException: already added: .... Conversion to Dalvik format failed with error 1 This I guess is coming because the android library(2.2) has been included twice in my apps( one for building my apps & another for building the open source). I dont know how to avoid this. Cleaning the project doesn't help. What i require is to use the open source and invoking it's activities in my apps without violating the constraints. If i can use the open source as bunch of .class files then great, or else any other way will do fine. Please look into it and let me know. Thanks

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  • Push notification is successfully sent, but the device does not receive (occasionally)

    - by ashiina
    I have been having a problem where some devices will not receive a push notification, since yesterday. The certificate / devicetoken seem to be correct, since the device used to successfully receive push notifications until yesterday. On the server-side, there are no errors or connection refusals, and the push notification seems to be successfully sent every time. But still, there are many occasions where the device does not correctly receive the push. Some surrounding information: I am doing this on the production environment. No errors / connection refusals on the server-side I am sending the exactly same JSON everytime. 2 of our devices are not receiving the push notification AT ALL since yesterday 1 of our device receives push notifications at a lower success rate (about 70%) than yesterday 1~2 of our devices still receive push notifications successfully even now. All of the above devices were able to receive push notifications properly on the production environment until yesterday. There is no difference in the server-side result for when the push is successful, and when the device doesn't receive it... Therefore it is virtually impossible to identify the problem. This is the server-side PHP code I am using: $ctx = stream_context_create(); stream_context_set_option($ctx, 'ssl', 'local_cert', $this->apnsData[$development]['certificate']); $fp = stream_socket_client($this->apnsData[$development]['ssl'], $error, $errorString, 100, (STREAM_CLIENT_C ONNECT|STREAM_CLIENT_PERSISTENT), $ctx); if(!$fp){ $this->_pushFailed($pid); $this->_triggerError("Failed to connect to APNS: {$error} {$errorString}."); } else { $msg = chr(0).pack("n",32).pack('H*',$token).pack("n",strlen($message)).$message; $fwrite = fwrite($fp, $msg); if(!$fwrite) { error_log("[APNS] push failed..."); $this->_pushFailed($pid); $this->_triggerError("Failed writing to stream.", E_USER_ERROR); } else { error_log("[APNS] push successful! ::: $token -> $message ($fwrite bytes)"); } } fclose($fp); The log tells me that the push was successful (Cutting out the token for privacy) : [Wed Dec 12 11:42:00 2012] [error] [client 10.161.6.177] [APNS] push successful! ::: aa4f******44 -> {"aps":{"alert":{"body":"\\u300casdfasdf\\u300d","action-loc-key":"OK"},"badge":4,"sound":"chime"}} (134 bytes) Is there any way I can get help on this problem? Or is there anybody who is having the same problem?? Please help! I am getting complaints from some users on this....

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  • dojo dgrid tree, subrows in wrong position

    - by Ventura
    I have a dgrid, working with tree column plugin. Every time that the user click on the tree, I call the server, catch the subrows(json) and bind it. But when it happens, these subrows are show in wrong position, like the image bellow. The most strange is when I change the pagination, after go back to first page, the subrows stay on the correct place. (please, tell me if is possible to understand my english, then I can try to improve the text) My dgrid code: var CustomGrid = declare([OnDemandGrid, Keyboard, Selection, Pagination]); var grid = new CustomGrid({ columns: [ selector({label: "#", disabled: function(object){ return object.type == 'DOCx'; }}, "radio"), {label:'Id', field:'id', sortable: false}, tree({label: "Title", field:"title", sortable: true, indentWidth:20, allowDuplicates:true}), //{label:'Title', field:'title', sortable: false}, {label:'Count', field:'count', sortable: false} ], store: this.memoryStore, collapseOnRefresh:true, pagingLinks: false, pagingTextBox: true, firstLastArrows: true, pageSizeOptions: [10, 15, 25], selectionMode: "single", // for Selection; only select a single row at a time cellNavigation: false // for Keyboard; allow only row-level keyboard navigation }, "grid"); My memory store: loadMemoryStore: function(items){ this.memoryStore = Observable(new Memory({ data: items, getChildren: function(parent, options){ return this.query({parent: parent.id}, options); }, mayHaveChildren: function(parent){ return (parent.count != 0) && (parent.type != 'DOC'); } })); }, This moment I am binding the subrows: success: function(data){ for(var i=0; i<data.report.length; i++){ this.memoryStore.put({id:data.report[i].id, title:data.report[i].created, type:'DOC', parent:this.designId}); } }, I was thinking, maybe every moment that I bind the subrows, I could do like a refresh on the grid, maybe works. I think that the pagination does the same thing. Thanks. edit: I forgot the question. Well, How can I correct this bug? If The refresh in dgrid works. How can I do it? Other thing that I was thinking, maybe my getChildren is wrong, but I could not identify it. thanks again.

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  • Why do we get a sudden spike in response times?

    - by Christian Hagelid
    We have an API that is implemented using ServiceStack which is hosted in IIS. While performing load testing of the API we discovered that the response times are good but that they deteriorate rapidly as soon as we hit about 3,500 concurrent users per server. We have two servers and when hitting them with 7,000 users the average response times sit below 500ms for all endpoints. The boxes are behind a load balancer so we get 3,500 concurrents per server. However as soon as we increase the number of total concurrent users we see a significant increase in response times. Increasing the concurrent users to 5,000 per server gives us an average response time per endpoint of around 7 seconds. The memory and CPU on the servers are quite low, both while the response times are good and when after they deteriorate. At peak with 10,000 concurrent users the CPU averages just below 50% and the RAM sits around 3-4 GB out of 16. This leaves us thinking that we are hitting some kind of limit somewhere. The below screenshot shows some key counters in perfmon during a load test with a total of 10,000 concurrent users. The highlighted counter is requests/second. To the right of the screenshot you can see the requests per second graph becoming really erratic. This is the main indicator for slow response times. As soon as we see this pattern we notice slow response times in the load test. How do we go about troubleshooting this performance issue? We are trying to identify if this is a coding issue or a configuration issue. Are there any settings in web.config or IIS that could explain this behaviour? The application pool is running .NET v4.0 and the IIS version is 7.5. The only change we have made from the default settings is to update the application pool Queue Length value from 1,000 to 5,000. We have also added the following config settings to the Aspnet.config file: <system.web> <applicationPool maxConcurrentRequestsPerCPU="5000" maxConcurrentThreadsPerCPU="0" requestQueueLimit="5000" /> </system.web> More details: The purpose of the API is to combine data from various external sources and return as JSON. It is currently using an InMemory cache implementation to cache individual external calls at the data layer. The first request to a resource will fetch all data required and any subsequent requests for the same resource will get results from the cache. We have a 'cache runner' that is implemented as a background process that updates the information in the cache at certain set intervals. We have added locking around the code that fetches data from the external resources. We have also implemented the services to fetch the data from the external sources in an asynchronous fashion so that the endpoint should only be as slow as the slowest external call (unless we have data in the cache of course). This is done using the System.Threading.Tasks.Task class. Could we be hitting a limitation in terms of number of threads available to the process?

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  • VPN in Ubuntu fails every time

    - by fazpas
    I am trying to setup a vpn connection in Ubuntu 10.04 to use the service from relakks.com I used the network manager to add the vpn connection and the settings are: Gateway: pptp.relakks.com Username: user Password: pwd IPv4 Settings: Automatic (VPN) Advanced: MSCHAP & MSCHAPv2 checked Use point-to-point encryption (security:default) Allow BSD data compression checked Allow deflate data compression checked Use TCP header compression checked The connection always fail, here is the syslog: Jun 27 20:11:56 desktop NetworkManager: <info> Starting VPN service 'org.freedesktop.NetworkManager.pptp'... Jun 27 20:11:56 desktop NetworkManager: <info> VPN service 'org.freedesktop.NetworkManager.pptp' started (org.freedesktop.NetworkManager.pptp), PID 2064 Jun 27 20:11:56 desktop NetworkManager: <info> VPN service 'org.freedesktop.NetworkManager.pptp' just appeared, activating connections Jun 27 20:11:56 desktop NetworkManager: <info> VPN plugin state changed: 3 Jun 27 20:11:56 desktop NetworkManager: <info> VPN connection 'Relakks' (Connect) reply received. Jun 27 20:11:56 desktop pppd[2067]: Plugin /usr/lib/pppd/2.4.5//nm-pptp-pppd-plugin.so loaded. Jun 27 20:11:56 desktop pppd[2067]: pppd 2.4.5 started by root, uid 0 Jun 27 20:11:56 desktop NetworkManager: SCPlugin-Ifupdown: devices added (path: /sys/devices/virtual/net/ppp1, iface: ppp1) Jun 27 20:11:56 desktop NetworkManager: SCPlugin-Ifupdown: device added (path: /sys/devices/virtual/net/ppp1, iface: ppp1): no ifupdown configuration found. Jun 27 20:11:56 desktop pppd[2067]: Using interface ppp1 Jun 27 20:11:56 desktop pppd[2067]: Connect: ppp1 <--> /dev/pts/0 Jun 27 20:11:56 desktop pptp[2071]: nm-pptp-service-2064 log[main:pptp.c:314]: The synchronous pptp option is NOT activated Jun 27 20:11:57 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 1 'Start-Control-Connection-Request' Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:739]: Received Start Control Connection Reply Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:773]: Client connection established. Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 7 'Outgoing-Call-Request' Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:858]: Received Outgoing Call Reply. Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:897]: Outgoing call established (call ID 0, peer's call ID 1024). Jun 27 20:11:59 desktop kernel: [ 56.564074] Inbound IN=ppp0 OUT= MAC= SRC=93.182.139.2 DST=186.110.76.26 LEN=61 TOS=0x00 PREC=0x00 TTL=52 ID=40460 DF PROTO=47 Jun 27 20:11:59 desktop kernel: [ 56.944054] Inbound IN=ppp0 OUT= MAC= SRC=93.182.139.2 DST=186.110.76.26 LEN=60 TOS=0x00 PREC=0x00 TTL=52 ID=40461 DF PROTO=47 Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[pptp_read_some:pptp_ctrl.c:544]: read returned zero, peer has closed Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[callmgr_main:pptp_callmgr.c:258]: Closing connection (shutdown) Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 12 'Call-Clear-Request' Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[pptp_read_some:pptp_ctrl.c:544]: read returned zero, peer has closed Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[call_callback:pptp_callmgr.c:79]: Closing connection (call state) Jun 27 20:11:59 desktop pppd[2067]: Modem hangup Jun 27 20:11:59 desktop pppd[2067]: Connection terminated. Jun 27 20:11:59 desktop NetworkManager: <info> VPN plugin failed: 1 Jun 27 20:11:59 desktop NetworkManager: SCPlugin-Ifupdown: devices removed (path: /sys/devices/virtual/net/ppp1, iface: ppp1) Jun 27 20:11:59 desktop pppd[2067]: Exit. Does someone can identify something in the syslog? I've been googling and reading about pptp but couldn't find anything about the error "read returned zero, peer has closed"

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  • I want to change DPI with Imagemagick without changing the actual byte-size of the image data

    - by user1694803
    I feel so horribly sorry that I have to ask this question here, but after hours of researching how to do an actually very simple task I'm still failing... In Gimp there is a very simple way to do what I want. I only have the German dialog installed but I'll try to translate it. I'm talking about going to "Picture-PrintingSize" and then adjusting the Values "X-Resolution" and "Y-Resolution" which are known to me as so called DPI values. You can also choose the format which by default is "Pixel/Inch". (In German the dialog is "Bild-Druckgröße" and there "X-Auflösung" and "Y-Auflösung") Ok, the values there are often "72" by default. When I change them to e.g. "300" this has the effect that the image stays the same on the computer, but if I print it, it will be smaller if you look at it, but all the details are still there, just smaller - it has a higher resolution on the printed paper (but smaller size... which is fine for me). I am often doing that when I am working with LaTeX, or to be exact with the command "pdflatex" on a recent Ubuntu-Machine. When I'm doing the above process with Gimp manually everything works just fine. The images will appear smaller in the resulting PDF but with high printing quality. What I am trying to do is to automate the process of going into Gimp and adjusting the DPI values. Since Imagemagick is known to be superb and I used it for many other tasks I tried to achieve my goal with this tool. But it does just not do what I want. After trying a lot of things I think this actually is be the command that should be my friend: convert input.png -density 300 output.png This should set the DPI to 300, as I can read everywhere in the web. It seems to work. When I check the file it stays the same. file input.png output.png input.png: PNG image data, 611 x 453, 8-bit grayscale, non-interlaced output.png: PNG image data, 611 x 453, 8-bit grayscale, non-interlaced When I use this command, it seems like it did what I wanted: identify -verbose output.png | grep 300 Resolution: 300x300 PNG:pHYs : x_res=300, y_res=300, units=0 (Funny enough, the same output comes for input.png which confuses me... so this might be the wrong parameters to watch?) But when I now render my TeX with "pdflatex" the image is still big and blurry. Also when I open the image with Gimp again the DPI values are set to "72" instead of "300". So there actually was no effect at all. Now what is the problem here. Am I getting something completely wrong? I can't be that wrong since everything works just fine with Gimp... Thanks for any help in this. I am also open to other automated solutions which are easily done on a Linux system...

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  • MySQL 5.5 (Percona) assertion failure log.. what would cause this?

    - by Tom Geee
    256GB, 64 Core , AMD running Ubuntu 12.04 with Percona MySQL 5.5.28. Below is the assertion failure. We just had a second assertion failure (different "in file", position, etc) while running a large set of inserts. After the first failure, MySQL restarted after a reboot only - after continuously looping on the same error after trying to recover. I decided to do a mysqlcheck with -o for optimize. Since these are all Innodb tables (very large tables, 60+GB) this would do an alter table on all tables. In the middle of this , the below assertion failure happened again: 121115 22:30:31 InnoDB: Assertion failure in thread 140086589445888 in file btr0pcur.c line 452 InnoDB: Failing assertion: btr_page_get_prev(next_page, mtr) == buf_block_get_page_no(btr_pcur_get_block(cursor)) InnoDB: We intentionally generate a memory trap. InnoDB: Submit a detailed bug report to http://bugs.mysql.com. InnoDB: If you get repeated assertion failures or crashes, even InnoDB: immediately after the mysqld startup, there may be InnoDB: corruption in the InnoDB tablespace. Please refer to InnoDB: http://dev.mysql.com/doc/refman/5.5/en/forcing-innodb-recovery.html InnoDB: about forcing recovery. 03:30:31 UTC - mysqld got signal 6 ; This could be because you hit a bug. It is also possible that this binary or one of the libraries it was linked against is corrupt, improperly built, or misconfigured. This error can also be caused by malfunctioning hardware. We will try our best to scrape up some info that will hopefully help diagnose the problem, but since we have already crashed, something is definitely wrong and this may fail. Please help us make Percona Server better by reporting any bugs at http://bugs.percona.com/ key_buffer_size=536870912 read_buffer_size=131072 max_used_connections=404 max_threads=500 thread_count=90 connection_count=90 It is possible that mysqld could use up to key_buffer_size + (read_buffer_size + sort_buffer_size)*max_threads = 1618416 K bytes of memory Hope that's ok; if not, decrease some variables in the equation. Thread pointer: 0x14edeb710 Attempting backtrace. You can use the following information to find out where mysqld died. If you see no messages after this, something went terribly wrong... stack_bottom = 7f687366ce80 thread_stack 0x30000 /usr/sbin/mysqld(my_print_stacktrace+0x2e)[0x7b52ee] /usr/sbin/mysqld(handle_fatal_signal+0x484)[0x68f024] /lib/x86_64-linux-gnu/libpthread.so.0(+0xfcb0)[0x7f9cbb23fcb0] /lib/x86_64-linux-gnu/libc.so.6(gsignal+0x35)[0x7f9cbaea6425] /lib/x86_64-linux-gnu/libc.so.6(abort+0x17b)[0x7f9cbaea9b8b] /usr/sbin/mysqld[0x858463] /usr/sbin/mysqld[0x804513] /usr/sbin/mysqld[0x808432] /usr/sbin/mysqld[0x7db8bf] /usr/sbin/mysqld(_Z13rr_sequentialP11READ_RECORD+0x1d)[0x755aed] /usr/sbin/mysqld(_Z17mysql_alter_tableP3THDPcS1_P24st_ha_create_informationP10TABLE_LISTP10Alter_infojP8st_orderb+0x216b)[0x60399b] /usr/sbin/mysqld(_Z20mysql_recreate_tableP3THDP10TABLE_LIST+0x166)[0x604bd6] /usr/sbin/mysqld[0x647da1] /usr/sbin/mysqld(_ZN24Optimize_table_statement7executeEP3THD+0xde)[0x64891e] /usr/sbin/mysqld(_Z21mysql_execute_commandP3THD+0x1168)[0x59b558] /usr/sbin/mysqld(_Z11mysql_parseP3THDPcjP12Parser_state+0x30c)[0x5a132c] /usr/sbin/mysqld(_Z16dispatch_command19enum_server_commandP3THDPcj+0x1620)[0x5a2a00] /usr/sbin/mysqld(_Z24do_handle_one_connectionP3THD+0x14f)[0x63ce6f] /usr/sbin/mysqld(handle_one_connection+0x51)[0x63cf31] /lib/x86_64-linux-gnu/libpthread.so.0(+0x7e9a)[0x7f9cbb237e9a] /lib/x86_64-linux-gnu/libc.so.6(clone+0x6d)[0x7f9cbaf63cbd] Trying to get some variables. Some pointers may be invalid and cause the dump to abort. Query (7f6300004b60): is an invalid pointer Connection ID (thread ID): 876 Status: NOT_KILLED You may download the Percona Server operations manual by visiting http://www.percona.com/software/percona-server/. You may find information in the manual which will help you identify the cause of the crash. 121115 22:31:07 [Note] Plugin 'FEDERATED' is disabled. 121115 22:31:07 InnoDB: The InnoDB memory heap is disabled 121115 22:31:07 InnoDB: Mutexes and rw_locks use GCC atomic builtins .. Then it recovered , without a reboot this time. from the log, what would cause this? I am currently running a dump to see if the problem resurfaces. edit: data partition is all in / since this is a hosted, defaulted file system unfortunately: Filesystem Size Used Avail Use% Mounted on /dev/vda3 742G 445G 260G 64% / udev 121G 4.0K 121G 1% /dev tmpfs 49G 248K 49G 1% /run none 5.0M 0 5.0M 0% /run/lock none 121G 0 121G 0% /run/shm /dev/vda1 99M 54M 40M 58% /boot my.cnf: [client] port = 3306 socket = /var/run/mysqld/mysqld.sock [mysqld_safe] socket = /var/run/mysqld/mysqld.sock nice = 0 [mysqld] skip-name-resolve innodb_file_per_table default_storage_engine=InnoDB user = mysql socket = /var/run/mysqld/mysqld.sock port = 3306 basedir = /usr datadir = /data/mysql tmpdir = /tmp skip-external-locking key_buffer = 512M max_allowed_packet = 128M thread_stack = 192K thread_cache_size = 64 myisam-recover = BACKUP max_connections = 500 table_cache = 812 table_definition_cache = 812 #query_cache_limit = 4M #query_cache_size = 512M join_buffer_size = 512K innodb_additional_mem_pool_size = 20M innodb_buffer_pool_size = 196G #innodb_file_io_threads = 4 #innodb_thread_concurrency = 12 innodb_flush_log_at_trx_commit = 1 innodb_log_buffer_size = 8M innodb_log_file_size = 1024M innodb_log_files_in_group = 2 innodb_max_dirty_pages_pct = 90 innodb_lock_wait_timeout = 120 log_error = /var/log/mysql/error.log long_query_time = 5 slow_query_log = 1 slow_query_log_file = /var/log/mysql/slowlog.log [mysqldump] quick quote-names max_allowed_packet = 16M [mysql] [isamchk] key_buffer = 16M

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  • VPN in Ubuntu fails every time

    - by fazpas
    I am trying to setup a vpn connection in Ubuntu 10.04 to use the service from relakks.com I used the network manager to add the vpn connection and the settings are: Gateway: pptp.relakks.com Username: user Password: pwd IPv4 Settings: Automatic (VPN) Advanced: MSCHAP & MSCHAPv2 checked Use point-to-point encryption (security:default) Allow BSD data compression checked Allow deflate data compression checked Use TCP header compression checked The connection always fail, here is the syslog: Jun 27 20:11:56 desktop NetworkManager: <info> Starting VPN service 'org.freedesktop.NetworkManager.pptp'... Jun 27 20:11:56 desktop NetworkManager: <info> VPN service 'org.freedesktop.NetworkManager.pptp' started (org.freedesktop.NetworkManager.pptp), PID 2064 Jun 27 20:11:56 desktop NetworkManager: <info> VPN service 'org.freedesktop.NetworkManager.pptp' just appeared, activating connections Jun 27 20:11:56 desktop NetworkManager: <info> VPN plugin state changed: 3 Jun 27 20:11:56 desktop NetworkManager: <info> VPN connection 'Relakks' (Connect) reply received. Jun 27 20:11:56 desktop pppd[2067]: Plugin /usr/lib/pppd/2.4.5//nm-pptp-pppd-plugin.so loaded. Jun 27 20:11:56 desktop pppd[2067]: pppd 2.4.5 started by root, uid 0 Jun 27 20:11:56 desktop NetworkManager: SCPlugin-Ifupdown: devices added (path: /sys/devices/virtual/net/ppp1, iface: ppp1) Jun 27 20:11:56 desktop NetworkManager: SCPlugin-Ifupdown: device added (path: /sys/devices/virtual/net/ppp1, iface: ppp1): no ifupdown configuration found. Jun 27 20:11:56 desktop pppd[2067]: Using interface ppp1 Jun 27 20:11:56 desktop pppd[2067]: Connect: ppp1 <--> /dev/pts/0 Jun 27 20:11:56 desktop pptp[2071]: nm-pptp-service-2064 log[main:pptp.c:314]: The synchronous pptp option is NOT activated Jun 27 20:11:57 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 1 'Start-Control-Connection-Request' Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:739]: Received Start Control Connection Reply Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:773]: Client connection established. Jun 27 20:11:58 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 7 'Outgoing-Call-Request' Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:858]: Received Outgoing Call Reply. Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_disp:pptp_ctrl.c:897]: Outgoing call established (call ID 0, peer's call ID 1024). Jun 27 20:11:59 desktop kernel: [ 56.564074] Inbound IN=ppp0 OUT= MAC= SRC=93.182.139.2 DST=186.110.76.26 LEN=61 TOS=0x00 PREC=0x00 TTL=52 ID=40460 DF PROTO=47 Jun 27 20:11:59 desktop kernel: [ 56.944054] Inbound IN=ppp0 OUT= MAC= SRC=93.182.139.2 DST=186.110.76.26 LEN=60 TOS=0x00 PREC=0x00 TTL=52 ID=40461 DF PROTO=47 Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[pptp_read_some:pptp_ctrl.c:544]: read returned zero, peer has closed Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[callmgr_main:pptp_callmgr.c:258]: Closing connection (shutdown) Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 12 'Call-Clear-Request' Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[pptp_read_some:pptp_ctrl.c:544]: read returned zero, peer has closed Jun 27 20:11:59 desktop pptp[2079]: nm-pptp-service-2064 log[call_callback:pptp_callmgr.c:79]: Closing connection (call state) Jun 27 20:11:59 desktop pppd[2067]: Modem hangup Jun 27 20:11:59 desktop pppd[2067]: Connection terminated. Jun 27 20:11:59 desktop NetworkManager: <info> VPN plugin failed: 1 Jun 27 20:11:59 desktop NetworkManager: SCPlugin-Ifupdown: devices removed (path: /sys/devices/virtual/net/ppp1, iface: ppp1) Jun 27 20:11:59 desktop pppd[2067]: Exit. Does someone can identify something in the syslog? I've been googling and reading about pptp but couldn't find anything about the error "read returned zero, peer has closed"

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  • Moved DNS and Email Hosting, Now Can't Send/Receive To/From Domains Hosted on Previous Host

    - by maxfinis
    Our company had 4 domains whose emails and DNS were hosted by one company, and then we moved the email and DNS hosting for 3 of the 4 domains to a new company. Now, the 3 domains that were moved can't send or receive emails to and from the one domain still left on the old server. All other email functions work fine for all 4 domains. There are no bouncebacks, error messages, or emails stuck in queue, and no evidence of these missing emails hitting the new servers. The new hosting company confirms that everything is fine on their end, and assures me that it's most likely an old zone file still remaining on the old nameserver, and so the emails sent from the old host is routed to what it believes is still the authoritative nameserver. Because the old zone file's MX records still contain the old resource, the requests never leave the old nameserver to go online to do a fresh search for the real (new) authoritative nameserver. The compounding problem is that the old company is rather inept and doesn't seem to have the technical expertise to identify the problem, much less fix it. (I know, I know.) Is the problem truly that this old zone file just needs to be deleted from the old company's nameserver? If so, what's the best way for me to describe this to them? If not, what do you think could be the issue? Any help is much appreciated. I'm not in IT, so all this is new to me. I know it seems weird for me (the client) to have to do this legwork, but I just want to get this resolved. Here's what I've done: Ran dig to verify that the old server's MX records still point to the old authoritative server, instead of going online to do a fresh search: ~$ dig @old.nameserver.com domainthatwasmoved.com mx ; << DiG 9.6.0-APPLE-P2 << @old.nameserver.com domainThatWasMoved.com mx ; (1 server found) ;; global options: +cmd ;; Got answer: ;; -HEADER<<- opcode: QUERY, status: NOERROR, id: 61227 ;; flags: qr aa rd ra; QUERY: 1, ANSWER: 1, AUTHORITY: 0, ADDITIONAL: 1 ;; QUESTION SECTION: ;domainthatwasmoved.com. IN MX ;; ANSWER SECTION: domainthatwasmoved.com. 3600 IN MX 10 mail.oldmailserver.com. ;; ADDITIONAL SECTION: mail.oldmailserver.com. 3600 IN A 65.198.191.5 ;; Query time: 29 msec ;; SERVER: 65.198.191.5#53(65.198.191.5) ;; WHEN: Sun Dec 26 16:59:22 2010 ;; MSG SIZE rcvd: 88 Ran dig to try to see where the new hosting company's servers look when emails are sent from the 3 domains that were moved, and got refused: ~$ dig @new.nameserver.net domainStillAtOldHost.com mx ; << DiG 9.6.0-APPLE-P2 << @new.nameserver.net domainStillAtOldHost.com mx ; (1 server found) ;; global options: +cmd ;; Got answer: ;; -HEADER<<- opcode: QUERY, status: REFUSED, id: 31599 ;; flags: qr rd; QUERY: 1, ANSWER: 0, AUTHORITY: 0, ADDITIONAL: 0 ;; WARNING: recursion requested but not available ;; QUESTION SECTION: ;domainStillAtOldHost.com. IN MX ;; Query time: 31 msec ;; SERVER: 216.201.128.10#53(216.201.128.10) ;; WHEN: Sun Dec 26 17:00:14 2010 ;; MSG SIZE rcvd: 34

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  • Why are USB 2.0 devices crashing my system?

    - by BenAlabaster
    Background on the machine I'm having a problem with: The machine was inherited and appears to be circa 2003 (there's a date stamp on the power supply which leads me to this conclusion). I've got it set up as a Skype terminal for my 2 year old to keep in touch with her grandparents and other members of the family - which everyone loves. It has a generic baby-ATX motherboard with no identifying markings. CPU-Z identifies the motherboard model as VT8601 but doesn't provide me with any manufacturer name. There's one stamp on the motherboard that says "Rev.B". On board it has 10/100 LAN, 2 x USB 1.0, VGA, PS/2 for KB and mouse, parallel port, 2 x serial ports, 2 x IDE, 1 x floppy, 2 x SDRAM slots, 1 x CPU housing that is seating a 1.3GHz Intel Celeron CPU, 3 x PCI, 1 x AGP - although you can only use 2 of the PCI slots if you use the AGP slot due to the physical layout of the board. It's got 768Mb PC133 SDRAM - 1 x 512Mb & 1 x 256Mb installed as well as a D-LINK WDA-2320 54G Wi-Fi network card and a generic USB 2.0 expansion board containing 3 x external + 1 x internal USB connectors. All this is sitting in a slimline case. I don't know the wattage of the PSU, but can post this later if this proves to be helpful. The motherboard is running a version of Award BIOS for which I don't have the version number to hand but can again post this later if it would be helpful. It has an 80Gb Western Digital hard drive freshly formatted and built with Windows XP Professional with Service Pack 3 and all current patches. In addition to Windows XP, the only other software it's running is Skype 4.1 (4.2 crashes the machine as soon as it starts up). It's got a Daytek MV150 15" touch screen running through the VGA and COM1 with the most current drivers from the Daytek website and the most current version of ELO-Touchsystems drivers for the touch component. The webcam is a Logitech Webcam C200 with the latest drivers from the Logitech website. The problem If I hook any USB 2.0 devices to this machine, it hangs the whole machine and I have to hard boot it to get it back up. Workarounds found I can plug the same devices into the on board USB 1.0 connectors and everything works fine, albeit at reduced performance. I've tried 3 different kinds of USB thumb drives, 3 different makes/models of webcams and my iPhone all with the same effect. They're recognized and don't hang the machine when I hook them to the USB 1.0 but if I hook them to the USB 2.0 ports, the machine hangs within a couple of seconds of recognizing the devices were connected. Attempted solutions I've tried disabling all the on board devices that I'm not using - such as the on board LAN, the second COM port, the AGP connector etc. through the BIOS in an (perhaps misguided or futile) attempt to reduce the power consumption... I don't think it had any effect but it didn't do any harm. I was wondering if the PSU wattage just isn't enough to drive the USB 2.0 devices; I've seen this suggested but haven't found any confirmation that this could really be an issue - nor have I found a way to work around this issue - if indeed it is one. Any ideas? The only thing I haven't done which I only just thought of while writing this essay is trying the USB 2.0 card in a different PCI slot, or re-ordering the wi-fi and USB cards in the slots... although I'm not sure if this will make any difference. I've installed the USB card in another machine and it works without issue, so it's not a problem with the USB card itself. Other thoughts Perhaps this is an incompatibility between the USB 2.0 card and the BIOS, would re-flashing the BIOS with a newer version help? Do I need to be able to identify the manufacturer of the motherboard in order to be able to find a BIOS edition specific for this motherboard or will any version of Award BIOS function in its place? Question Does anyone have any ideas that could help me get my USB 2.0 devices hooked up to this machine?

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  • Configuring nginx server to handle requests from multiple domains

    - by KillABug
    Use Case:- I am working on a web application which allows to create HTML templates and publish them on amazon S3.Now to publish the websites I use nginx as a proxy server. What the proxy server does is,when a user enters the website URL,I want to identify how to check if the request comes from my application i.e app.mysite.com(This won't change) and route it to apache for regular access,if its coming from some other domain like a regular URL www.mysite.com(This needs to be handled dynamically.Can be random) it goes to the S3 bucket that hosts the template. My current configuration is: user nginx; worker_processes 1; error_log /var/log/nginx/error.log; pid /var/run/nginx.pid; events { worker_connections 1024; } http { include /etc/nginx/mime.types; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log /var/log/nginx/access.log main; charset utf-8; keepalive_timeout 65; server_tokens off; sendfile on; tcp_nopush on; tcp_nodelay off; Default Server Block to catch undefined host names server { listen 80; server_name app.mysite.com; access_log off; error_log off; location / { proxy_pass http://127.0.0.1:8080; proxy_set_header X-Real-IP $remote_addr; proxy_set_header Host $host; proxy_redirect off; proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_connect_timeout 90; proxy_send_timeout 90; proxy_read_timeout 90; client_max_body_size 10m; client_body_buffer_size 128k; proxy_buffer_size 4k; proxy_buffers 4 32k; proxy_busy_buffers_size 64k; } } } Load all the sites include /etc/nginx/conf.d/*.conf; Updates as I was not clear enough :- My question is how can I handle both the domains in the config file.My nginx is a proxy server on port 80 on an EC2 instance.This also hosts my application that runs on apache on a differnet port.So any request coming for my application will come from a domain app.mysite.com and I also want to proxy the hosted templates on S3 which are inside a bucket say sites.mysite.com/coolsite.com/index.html.So if someone hits coolsite.com I want to proxy it to the folder sites.mysite.com/coolsite.com/index.html and not to app.syartee.com.Hope I am clear The other server block: # Server for S3 server { # Listen on port 80 for all IPs associated with your machine listen 80; # Catch all other server names server_name _; //I want it to handle other domains then app.mysite.com # This code gets the host without www. in front and places it inside # the $host_without_www variable # If someone requests www.coolsite.com, then $host_without_www will have the value coolsite.com set $host_without_www $host; if ($host ~* www\.(.*)) { set $host_without_www $1; } location / { # This code rewrites the original request, and adds the host without www in front # E.g. if someone requests # /directory/file.ext?param=value # from the coolsite.com site the request is rewritten to # /coolsite.com/directory/file.ext?param=value set $foo 'http://sites.mysite.com'; # echo "$foo"; rewrite ^(.*)$ $foo/$host_without_www$1 break; # The rewritten request is passed to S3 proxy_pass http://sites.mysite.com; include /etc/nginx/proxy_params; } } Also I understand I will have to make the DNS changes in the cname of the domain.I guess I will have to add app.mysite.com under the CNAME of the template domain name?Please correct if wrong. Thank you for your time

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  • Issue in nginx proxying to apache

    - by Luis Masuelli
    My current nginx configuration is as follows: specific configuration for (currently two) domains: server { listen 443 ssl; server_name studiotv.service.tebusco.lan phpmyadmin.service.tebusco.lan; ssl_certificate /home/administrador/nginx-confs/ssl/service.tebusco.lan.crt; ssl_certificate_key /home/administrador/nginx-confs/ssl/service.tebusco.lan.key; location / { proxy_pass http://127.0.0.1:8180; proxy_set_header Host $http_host:8180; } } default configuration for unmatched ssl connections: server { listen 443 default ssl; ssl_certificate /home/administrador/nginx-confs/ssl/service.tebusco.lan.crt; ssl_certificate_key /home/administrador/nginx-confs/ssl/service.tebusco.lan.key; location / { return 403; } } http configuration: server { listen 80; rewrite ^ https://$host$request_uri? permanent; } The intention is clear: Redirect http traffic to https. Proxy each https:// call from phpmyadmin.service.tebusco.lan and studiotv.service.tebusco.lan to apache2. This includes passing a host header, which is detected. Each unmatched ssl connection must return a 403 in nginx. Does not even reach apache2. In the apache2 side of the life, I have a default site, and a non-default site which will match studiotv.service.tebusco.lan: 000-default.conf file (available and enabled): <VirtualHost 127.0.0.1:8180> # The ServerName directive sets the request scheme, hostname and port that # the server uses to identify itself. This is used when creating # redirection URLs. In the context of virtual hosts, the ServerName # specifies what hostname must appear in the request's Host: header to # match this virtual host. For the default virtual host (this file) this # value is not decisive as it is used as a last resort host regardless. # However, you must set it for any further virtual host explicitly. ServerName localhost ServerAdmin webmaster@localhost DocumentRoot /var/www/html <Directory /var/www/html> Order deny,allow Require all granted </Directory> </VirtualHost> # vim: syntax=apache ts=4 sw=4 sts=4 sr noet studiotv.conf file (available and enabled): <VirtualHost *:8180> ServerName studiotv.service.tebusco.lan ServerAdmin [email protected] DocumentRoot /var/www/studiotv <Directory /var/www/studiotv/> Options -Indexes +FollowSymLinks AllowOverride None Order deny,allow Allow from all Require all granted </Directory> # Available loglevels: trace8, ..., trace1, debug, info, notice, warn, # error, crit, alert, emerg. # It is also possible to configure the loglevel for particular # modules, e.g. #LogLevel info ssl:warn # No usamos ${APACHE_LOG_DIR} sino en su lugar /var/log/<host> ErrorLog /var/log/apache2/studiotv/error.log CustomLog /var/log/apache2/studiotv/access.log combined </VirtualHost> # vim: syntax=apache ts=4 sw=4 sts=4 sr noet However, when I hit the browser with http://studiotv.service.tebusco.lan, the default php page is shown instead. Question: What am I missing? (apache 2.4.7, nginx 1.6.0, ubuntu server 14.04).

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  • How to save a ntfs partition which suddenly became empty

    - by SteveO
    One ntfs partition of my laptop was suddenly wiped out without any notice to me, when I rebooted from Windows 7 to Ubuntu 12.04 today. I am in need of help to save my files on that partition, which are important and unfortunately haven't been backed up yet. My laptop has two operating systems: Windows 7 and Ubuntu 12.04. with a ntfs partition shared between the two operating systems for storing some data files (109GB, about 97%of which has been used). I have almost always been using Ubuntu, but today I happened to have to work under Windows. Following is a record of what happened in the time order, numbering according to which operating system I was in at each stage. When I started into Windows 7, right before being able to log in, it took a while and two reboots to configure the Windows. I thought it was normal, since last time when I was using Windows two weeks ago, it took very long and several reboots to update Windows, since the last time I used Windows before then was in November last year. Then after finally being able to log in Windows 7, I installed Libre Office, MathType (I got it from http://dl.portablesoft.org/down/?id=2515, which I originally thought was a trial version, but later I learned was a cracked version and felt wrong. I made a copy of it at dropbox http://dl.dropbox.com/u/13029929/MathType_6.8_PortableSoft.rar, not for distributing it but to list it there just in case it will help to identify the problem), and MikTex. I then edited some .doc files in the ntfs partition under both Microsoft Office with MathType, and Libre Office. When I finished working under Windows and rebooted into Ubuntu, Ubuntu did some filesystem checking and reported that the ntfs partition was not able to be mounted. Then I rebooted again into Windows, and found that the ntfs partition had been emptied, i.e. all the data files were gone, and only one system file bootsqm.dat and one system directory System Volume Information were there, with their last updated time being the time when I first rebooted from Windows to Ubuntu (in fact, it is 4 hours in advanced than the actual time of that rebooting , see immediately below) Also I noticed that the time shown by Windows is not correct for my time zone (UTC-05:00) Eastern Time (US & Canada)), which is 4 hours in advance than the correct time (my current time is 3am, but the computer shows 7am). Same things happened when I rebooted into Ubuntu again: the ntfs has been emptied and left with only one Windows system file bootsqm.dat and one Windows system directory System Volume Information. the time shown by Ubuntu is 4 hours in advance than the correct time. I wonder what I can do to retrieve my data files back on the ntfs partition? If I am not able to do it myself, will some professionals be able to help me out? Thanks a lot! PS: I didn't think I did any thing that required emptying that partition. But there were quite some works I did during that stage right before the reboot from Windows to Ubuntu when the problem occured. Did I make any mis-operation?

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  • Enabling Kerberos Authentication for Reporting Services

    - by robcarrol
    Recently, I’ve helped several customers with Kerberos authentication problems with Reporting Services and Analysis Services, so I’ve decided to write this blog post and pull together some useful resources in one place (there are 2 whitepapers in particular that I found invaluable configuring Kerberos authentication, and these can be found in the references section at the bottom of this post). In most of these cases, the problem has manifested itself with the Login failed for User ‘NT Authority\Anonymous’ (“double-hop”) error. By default, Reporting Services uses Windows Integrated Authentication, which includes the Kerberos and NTLM protocols for network authentication. Additionally, Windows Integrated Authentication includes the negotiate security header, which prompts the client to select Kerberos or NTLM for authentication. The client can access reports which have the appropriate permissions by using Kerberos for authentication. Servers that use Kerberos authentication can impersonate those clients and use their security context to access network resources. You can configure Reporting Services to use both Kerberos and NTLM authentication; however this may lead to a failure to authenticate. With negotiate, if Kerberos cannot be used, the authentication method will default to NTLM. When negotiate is enabled, the Kerberos protocol is always used except when: Clients/servers that are involved in the authentication process cannot use Kerberos. The client does not provide the information necessary to use Kerberos. An in-depth discussion of Kerberos authentication is beyond the scope of this post, however when users execute reports that are configured to use Windows Integrated Authentication, their logon credentials are passed from the report server to the server hosting the data source. Delegation needs to be set on the report server and Service Principle Names (SPNs) set for the relevant services. When a user processes a report, the request must go through a Web server on its way to a database server for processing. Kerberos authentication enables the Web server to request a service ticket from the domain controller; impersonate the client when passing the request to the database server; and then restrict the request based on the user’s permissions. Each time a server is required to pass the request to another server, the same process must be used. Kerberos authentication is supported in both native and SharePoint integrated mode, but I’ll focus on native mode for the purpose of this post (I’ll explain configuring SharePoint integrated mode and Kerberos authentication in a future post). Configuring Kerberos avoids the authentication failures due to double-hop issues. These double-hop errors occur when a users windows domain credentials can’t be passed to another server to complete the user’s request. In the case of my customers, users were executing Reporting Services reports that were configured to query Analysis Services cubes on a separate machine using Windows Integrated security. The double-hop issue occurs as NTLM credentials are valid for only one network hop, subsequent hops result in anonymous authentication. The client attempts to connect to the report server by making a request from a browser (or some other application), and the connection process begins with authentication. With NTLM authentication, client credentials are presented to Computer 2. However Computer 2 can’t use the same credentials to access Computer 3 (so we get the Anonymous login error). To access Computer 3 it is necessary to configure the connection string with stored credentials, which is what a number of customers I have worked with have done to workaround the double-hop authentication error. However, to get the benefits of Windows Integrated security, a better solution is to enable Kerberos authentication. Again, the connection process begins with authentication. With Kerberos authentication, the client and the server must demonstrate to one another that they are genuine, at which point authentication is successful and a secure client/server session is established. In the illustration above, the tiers represent the following: Client tier (computer 1): The client computer from which an application makes a request. Middle tier (computer 2): The Web server or farm where the client’s request is directed. Both the SharePoint and Reporting Services server(s) comprise the middle tier (but we’re only concentrating on native deployments just now). Back end tier (computer 3): The Database/Analysis Services server/Cluster where the requested data is stored. In order to enable Kerberos authentication for Reporting Services it’s necessary to configure the relevant SPNs, configure trust for delegation for server accounts, configure Kerberos with full delegation and configure the authentication types for Reporting Services. Service Principle Names (SPNs) are unique identifiers for services and identify the account’s type of service. If an SPN is not configured for a service, a client account will be unable to authenticate to the servers using Kerberos. You need to be a domain administrator to add an SPN, which can be added using the SetSPN utility. For Reporting Services in native mode, the following SPNs need to be registered --SQL Server Service SETSPN -S mssqlsvc/servername:1433 Domain\SQL For named instances, or if the default instance is running under a different port, then the specific port number should be used. --Reporting Services Service SETSPN -S http/servername Domain\SSRS SETSPN -S http/servername.domain.com Domain\SSRS The SPN should be set for the NETBIOS name of the server and the FQDN. If you access the reports using a host header or DNS alias, then that should also be registered SETSPN -S http/www.reports.com Domain\SSRS --Analysis Services Service SETSPN -S msolapsvc.3/servername Domain\SSAS Next, you need to configure trust for delegation, which refers to enabling a computer to impersonate an authenticated user to services on another computer: Location Description Client 1. The requesting application must support the Kerberos authentication protocol. 2. The user account making the request must be configured on the domain controller. Confirm that the following option is not selected: Account is sensitive and cannot be delegated. Servers 1. The service accounts must be trusted for delegation on the domain controller. 2. The service accounts must have SPNs registered on the domain controller. If the service account is a domain user account, the domain administrator must register the SPNs. In Active Directory Users and Computers, verify that the domain user accounts used to access reports have been configured for delegation (the ‘Account is sensitive and cannot be delegated’ option should not be selected): We then need to configure the Reporting Services service account and computer to use Kerberos with full delegation:   We also need to do the same for the SQL Server or Analysis Services service accounts and computers (depending on what type of data source you are connecting to in your reports). Finally, and this is the part that sometimes gets over-looked, we need to configure the authentication type correctly for reporting services to use Kerberos authentication. This is configured in the Authentication section of the RSReportServer.config file on the report server. <Authentication> <AuthenticationTypes>           <RSWindowsNegotiate/> </AuthenticationTypes> <EnableAuthPersistence>true</EnableAuthPersistence> </Authentication> This will enable Kerberos authentication for Internet Explorer. For other browsers, see the link below. The report server instance must be restarted for these changes to take effect. Once these changes have been made, all that’s left to do is test to make sure Kerberos authentication is working properly by running a report from report manager that is configured to use Windows Integrated authentication (either connecting to Analysis Services or SQL Server back-end). Resources: Manage Kerberos Authentication Issues in a Reporting Services Environment http://download.microsoft.com/download/B/E/1/BE1AABB3-6ED8-4C3C-AF91-448AB733B1AF/SSRSKerberos.docx Configuring Kerberos Authentication for Microsoft SharePoint 2010 Products http://www.microsoft.com/download/en/details.aspx?displaylang=en&id=23176 How to: Configure Windows Authentication in Reporting Services http://msdn.microsoft.com/en-us/library/cc281253.aspx RSReportServer Configuration File http://msdn.microsoft.com/en-us/library/ms157273.aspx#Authentication Planning for Browser Support http://msdn.microsoft.com/en-us/library/ms156511.aspx

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  • Converting a Visual Studio 2003 Web Project to a Visual Studio 2008 Web Application Project

    - by navaneeth
    This walkthrough describes how to convert a Visual Studio .NET 2002 or Visual Studio .NET 2003 Web project to a Visual Studio 2008 Web application project. The Visual Studio 2008 Web application project model is like the Visual Studio 2005 Web application project model. Therefore, the conversion processes are similar. For more information about Web application projects, see ASP.NET Web Application Projects. You can also convert from a Visual Studio .NET Web project to a Visual Studio 2008 Web site project. However, conversion to a Web application project is the approach that is supported, and gives you the convenience of tools to help with the conversion. For example, when you convert to a Visual Studio 2008 Web application project, you can use the Visual Studio Conversion Wizard to automate part of the process. For information about how to convert a Visual Studio .NET Web project to a Visual Studio 2008 Web site, see Common Web Project Conversion Issues and Solutions. There are two parts involved in converting a Visual Studio 2002 or 2003 Web project to a Visual Studio 2008 Web application project. The parts are as follows: Converting the project. You can use the Visual Studio Conversion Wizard for the initial conversion of the project and Web.config files. You can later use the Convert To Web Application command to update the project's files and structure. Upgrading the .NET Framework version of the project. You must upgrade the project's .NET Framework version to either .NET Framework 2.0 SP1 or to .NET Framework 3.5. This .NET Framework version upgrade is required because Visual Studio 2008 cannot target earlier versions of the .NET Framework. You can perform this upgrade during the project conversion, by using the Conversion Wizard. Alternatively, you can upgrade the .NET Framework version after you convert the project.   NoteYou can change a project's .NET Framework version manually. To do so, in Visual Studio open the property pages for the project, click the Application tab, and then select a new version from the Target Framework list. This walkthrough illustrates the following tasks: Opening the Visual Studio .NET project in Visual Studio 2008 and creating a backup of the project files. Upgrading the .NET Framework version that the project targets. Converting the project file and the Web.config file. Converting ASP.NET code files. Testing the converted project. Prerequisites    To complete this walkthrough, you will need: Visual Studio 2008. A Web site project that was created in Visual Studio .NET version 2002 or 2003 that compiles and runs without errors. Converting the Project and Upgrading the .NET Framework Version    To begin, you open the project in Visual Studio 2008, which starts the conversion. It offers you an opportunity to back up the project before converting it. NoteIt is strongly recommended that you back up the project. The conversion works on the original project files, which cannot be recovered if the conversion is not successful.To convert the project and back up the files In Visual Studio 2008, in the File menu, click Open and then click Project. The Open Project dialog box is displayed. Browse to the folder that contains the project or solution file for the Visual Studio .NET project, select the file, and then click Open. NoteMake sure that you open the project by using the Open Project command. If you use the Open Web Site command, the project will be converted to the Web site project format.The Conversion Wizard opens and prompts you to create a backup before converting the project. To create the backup, click Yes. Click Browse, select the folder in which the backup should be created, and then click Next. Click Finish. The backup starts. NoteThere might be significant delays as the Conversion Wizard copies files, with no updates or progress indicated. Wait until the process finishes before you continue.When the conversion finishes, the wizard prompts you to upgrade the targeted version of the .NET Framework for the project. To upgrade to the .NET Framework 3.5, click Yes. To upgrade the project to target the .NET Framework 2.0 SP1, click No. It is recommended that you leave the check box selected that asks whether you want to upgrade all Webs in the solution. If you upgrade to .NET Framework 3.5, the project's Web.config file is modified at the same time as the project file. When the upgrade and conversion have finished, a message is displayed that indicates that you have completed the first step in converting your project. Click OK. The wizard displays status information about the conversion. Click Close. Testing the Converted Project    After the conversion has finished, you can test the project to make sure that it runs. This will also help you identify code in the project that must be updated. To verify that the project runs If you know about changes that are required for the code to run with the new version of the .NET Framework, make those changes. In the Build menu, click Build. Any missing references or other compilation issues in the project are displayed in the Error List window. The most likely issues are missing assembly references or issues with dynamically generated types. In Solution Explorer, right-click the Web page that will be used to launch the application, and then click Set as Start Page. On the Debug menu, click Start Debugging. If debugging is not enabled, the Debugging Not Enabled dialog box is displayed. Select the option to add a Web.config file that has debugging enabled, and then click OK. Verify that the converted project runs as expected. Do not continue with the conversion process until all build and run-time errors are resolved. Converting ASP.NET Code Files    ASP.NET Web page files and user-control files in Visual Studio 2008 that use the code-behind model have an associated designer file. The files that you just converted will have an associated code-behind file, but no designer file. Therefore, the next step is to generate designer files. NoteOnly ASP.NET Web pages and user controls that have their code in a separate code file require a separate designer file. For pages that have inline code and no associated code file, no designer file will be generated.To convert ASP.NET code files In Solution Explorer, right-click the project node, and then click Convert To Web Application. The files are converted. Verify that the converted code files have a code file and a designer file. Build and run the project to verify the results of the conversion.

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  • SQL SERVER – How to Recover SQL Database Data Deleted by Accident

    - by Pinal Dave
    In Repair a SQL Server database using a transaction log explorer, I showed how to use ApexSQL Log, a SQL Server transaction log viewer, to recover a SQL Server database after a disaster. In this blog, I’ll show you how to use another SQL Server disaster recovery tool from ApexSQL in a situation when data is accidentally deleted. You can download ApexSQL Recover here, install, and play along. With a good SQL Server disaster recovery strategy, data recovery is not a problem. You have a reliable full database backup with valid data, a full database backup and subsequent differential database backups, or a full database backup and a chain of transaction log backups. But not all situations are ideal. Here we’ll address some sub-optimal scenarios, where you can still successfully recover data. If you have only a full database backup This is the least optimal SQL Server disaster recovery strategy, as it doesn’t ensure minimal data loss. For example, data was deleted on Wednesday. Your last full database backup was created on Sunday, three days before the records were deleted. By using the full database backup created on Sunday, you will be able to recover SQL database records that existed in the table on Sunday. If there were any records inserted into the table on Monday or Tuesday, they will be lost forever. The same goes for records modified in this period. This method will not bring back modified records, only the old records that existed on Sunday. If you restore this full database backup, all your changes (intentional and accidental) will be lost and the database will be reverted to the state it had on Sunday. What you have to do is compare the records that were in the table on Sunday to the records on Wednesday, create a synchronization script, and execute it against the Wednesday database. If you have a full database backup followed by differential database backups Let’s say the situation is the same as in the example above, only you create a differential database backup every night. Use the full database backup created on Sunday, and the last differential database backup (created on Tuesday). In this scenario, you will lose only the data inserted and updated after the differential backup created on Tuesday. If you have a full database backup and a chain of transaction log backups This is the SQL Server disaster recovery strategy that provides minimal data loss. With a full chain of transaction logs, you can recover the SQL database to an exact point in time. To provide optimal results, you have to know exactly when the records were deleted, because restoring to a later point will not bring back the records. This method requires restoring the full database backup first. If you have any differential log backup created after the last full database backup, restore the most recent one. Then, restore transaction log backups, one by one, it the order they were created starting with the first created after the restored differential database backup. Now, the table will be in the state before the records were deleted. You have to identify the deleted records, script them and run the script against the original database. Although this method is reliable, it is time-consuming and requires a lot of space on disk. How to easily recover deleted records? The following solution enables you to recover SQL database records even if you have no full or differential database backups and no transaction log backups. To understand how ApexSQL Recover works, I’ll explain what happens when table data is deleted. Table data is stored in data pages. When you delete table records, they are not immediately deleted from the data pages, but marked to be overwritten by new records. Such records are not shown as existing anymore, but ApexSQL Recover can read them and create undo script for them. How long will deleted records stay in the MDF file? It depends on many factors, as time passes it’s less likely that the records will not be overwritten. The more transactions occur after the deletion, the more chances the records will be overwritten and permanently lost. Therefore, it’s recommended to create a copy of the database MDF and LDF files immediately (if you cannot take your database offline until the issue is solved) and run ApexSQL Recover on them. Note that a full database backup will not help here, as the records marked for overwriting are not included in the backup. First, I’ll delete some records from the Person.EmailAddress table in the AdventureWorks database.   I can delete these records in SQL Server Management Studio, or execute a script such as DELETE FROM Person.EmailAddress WHERE BusinessEntityID BETWEEN 70 AND 80 Then, I’ll start ApexSQL Recover and select From DELETE operation in the Recovery tab.   In the Select the database to recover step, first select the SQL Server instance. If it’s not shown in the drop-down list, click the Server icon right to the Server drop-down list and browse for the SQL Server instance, or type the instance name manually. Specify the authentication type and select the database in the Database drop-down list.   In the next step, you’re prompted to add additional data sources. As this can be a tricky step, especially for new users, ApexSQL Recover offers help via the Help me decide option.   The Help me decide option guides you through a series of questions about the database transaction log and advises what files to add. If you know that you have no transaction log backups or detached transaction logs, or the online transaction log file has been truncated after the data was deleted, select No additional transaction logs are available. If you know that you have transaction log backups that contain the delete transactions you want to recover, click Add transaction logs. The online transaction log is listed and selected automatically.   Click Add if to add transaction log backups. It would be best if you have a full transaction log chain, as explained above. The next step for this option is to specify the time range.   Selecting a small time range for the time of deletion will create the recovery script just for the accidentally deleted records. A wide time range might script the records deleted on purpose, and you don’t want that. If needed, you can check the script generated and manually remove such records. After that, for all data sources options, the next step is to select the tables. Be careful here, if you deleted some data from other tables on purpose, and don’t want to recover them, don’t select all tables, as ApexSQL Recover will create the INSERT script for them too.   The next step offers two options: to create a recovery script that will insert the deleted records back into the Person.EmailAddress table, or to create a new database, create the Person.EmailAddress table in it, and insert the deleted records. I’ll select the first one.   The recovery process is completed and 11 records are found and scripted, as expected.   To see the script, click View script. ApexSQL Recover has its own script editor, where you can review, modify, and execute the recovery script. The insert into statements look like: INSERT INTO Person.EmailAddress( BusinessEntityID, EmailAddressID, EmailAddress, rowguid, ModifiedDate) VALUES( 70, 70, N'[email protected]' COLLATE SQL_Latin1_General_CP1_CI_AS, 'd62c5b4e-c91f-403f-b630-7b7e0fda70ce', '20030109 00:00:00.000' ); To execute the script, click Execute in the menu.   If you want to check whether the records are really back, execute SELECT * FROM Person.EmailAddress WHERE BusinessEntityID BETWEEN 70 AND 80 As shown, ApexSQL Recover recovers SQL database data after accidental deletes even without the database backup that contains the deleted data and relevant transaction log backups. ApexSQL Recover reads the deleted data from the database data file, so this method can be used even for databases in the Simple recovery model. Besides recovering SQL database records from a DELETE statement, ApexSQL Recover can help when the records are lost due to a DROP TABLE, or TRUNCATE statement, as well as repair a corrupted MDF file that cannot be attached to as SQL Server instance. You can find more information about how to recover SQL database lost data and repair a SQL Server database on ApexSQL Solution center. There are solutions for various situations when data needs to be recovered. Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: PostADay, SQL, SQL Authority, SQL Backup and Restore, SQL Query, SQL Server, SQL Tips and Tricks, T SQL

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