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  • [GEEK SCHOOL] Network Security 1: Securing User Accounts and Passwords in Windows

    - by Matt Klein
    This How-To Geek School class is intended for people who want to learn more about security when using Windows operating systems. You will learn many principles that will help you have a more secure computing experience and will get the chance to use all the important security tools and features that are bundled with Windows. Obviously, we will share everything you need to know about using them effectively. In this first lesson, we will talk about password security; the different ways of logging into Windows and how secure they are. In the proceeding lesson, we will explain where Windows stores all the user names and passwords you enter while working in this operating systems, how safe they are, and how to manage this data. Moving on in the series, we will talk about User Account Control, its role in improving the security of your system, and how to use Windows Defender in order to protect your system from malware. Then, we will talk about the Windows Firewall, how to use it in order to manage the apps that get access to the network and the Internet, and how to create your own filtering rules. After that, we will discuss the SmartScreen Filter – a security feature that gets more and more attention from Microsoft and is now widely used in its Windows 8.x operating systems. Moving on, we will discuss ways to keep your software and apps up-to-date, why this is important and which tools you can use to automate this process as much as possible. Last but not least, we will discuss the Action Center and its role in keeping you informed about what’s going on with your system and share several tips and tricks about how to stay safe when using your computer and the Internet. Let’s get started by discussing everyone’s favorite subject: passwords. The Types of Passwords Found in Windows In Windows 7, you have only local user accounts, which may or may not have a password. For example, you can easily set a blank password for any user account, even if that one is an administrator. The only exception to this rule are business networks where domain policies force all user accounts to use a non-blank password. In Windows 8.x, you have both local accounts and Microsoft accounts. If you would like to learn more about them, don’t hesitate to read the lesson on User Accounts, Groups, Permissions & Their Role in Sharing, in our Windows Networking series. Microsoft accounts are obliged to use a non-blank password due to the fact that a Microsoft account gives you access to Microsoft services. Using a blank password would mean exposing yourself to lots of problems. Local accounts in Windows 8.1 however, can use a blank password. On top of traditional passwords, any user account can create and use a 4-digit PIN or a picture password. These concepts were introduced by Microsoft to speed up the sign in process for the Windows 8.x operating system. However, they do not replace the use of a traditional password and can be used only in conjunction with a traditional user account password. Another type of password that you encounter in Windows operating systems is the Homegroup password. In a typical home network, users can use the Homegroup to easily share resources. A Homegroup can be joined by a Windows device only by using the Homegroup password. If you would like to learn more about the Homegroup and how to use it for network sharing, don’t hesitate to read our Windows Networking series. What to Keep in Mind When Creating Passwords, PINs and Picture Passwords When creating passwords, a PIN, or a picture password for your user account, we would like you keep in mind the following recommendations: Do not use blank passwords, even on the desktop computers in your home. You never know who may gain unwanted access to them. Also, malware can run more easily as administrator because you do not have a password. Trading your security for convenience when logging in is never a good idea. When creating a password, make it at least eight characters long. Make sure that it includes a random mix of upper and lowercase letters, numbers, and symbols. Ideally, it should not be related in any way to your name, username, or company name. Make sure that your passwords do not include complete words from any dictionary. Dictionaries are the first thing crackers use to hack passwords. Do not use the same password for more than one account. All of your passwords should be unique and you should use a system like LastPass, KeePass, Roboform or something similar to keep track of them. When creating a PIN use four different digits to make things slightly harder to crack. When creating a picture password, pick a photo that has at least 10 “points of interests”. Points of interests are areas that serve as a landmark for your gestures. Use a random mixture of gesture types and sequence and make sure that you do not repeat the same gesture twice. Be aware that smudges on the screen could potentially reveal your gestures to others. The Security of Your Password vs. the PIN and the Picture Password Any kind of password can be cracked with enough effort and the appropriate tools. There is no such thing as a completely secure password. However, passwords created using only a few security principles are much harder to crack than others. If you respect the recommendations shared in the previous section of this lesson, you will end up having reasonably secure passwords. Out of all the log in methods in Windows 8.x, the PIN is the easiest to brute force because PINs are restricted to four digits and there are only 10,000 possible unique combinations available. The picture password is more secure than the PIN because it provides many more opportunities for creating unique combinations of gestures. Microsoft have compared the two login options from a security perspective in this post: Signing in with a picture password. In order to discourage brute force attacks against picture passwords and PINs, Windows defaults to your traditional text password after five failed attempts. The PIN and the picture password function only as alternative login methods to Windows 8.x. Therefore, if someone cracks them, he or she doesn’t have access to your user account password. However, that person can use all the apps installed on your Windows 8.x device, access your files, data, and so on. How to Create a PIN in Windows 8.x If you log in to a Windows 8.x device with a user account that has a non-blank password, then you can create a 4-digit PIN for it, to use it as a complementary login method. In order to create one, you need to go to “PC Settings”. If you don’t know how, then press Windows + C on your keyboard or flick from the right edge of the screen, on a touch-enabled device, then press “Settings”. The Settings charm is now open. Click or tap the link that says “Change PC settings”, on the bottom of the charm. In PC settings, go to Accounts and then to “Sign-in options”. Here you will find all the necessary options for changing your existing password, creating a PIN, or a picture password. To create a PIN, press the “Add” button in the PIN section. The “Create a PIN” wizard is started and you are asked to enter the password of your user account. Type it and press “OK”. Now you are asked to enter a 4-digit pin in the “Enter PIN” and “Confirm PIN” fields. The PIN has been created and you can now use it to log in to Windows. How to Create a Picture Password in Windows 8.x If you log in to a Windows 8.x device with a user account that has a non-blank password, then you can also create a picture password and use it as a complementary login method. In order to create one, you need to go to “PC settings”. In PC Settings, go to Accounts and then to “Sign-in options”. Here you will find all the necessary options for changing your existing password, creating a PIN, or a picture password. To create a picture password, press the “Add” button in the “Picture password” section. The “Create a picture password” wizard is started and you are asked to enter the password of your user account. You are shown a guide on how the picture password works. Take a few seconds to watch it and learn the gestures that can be used for your picture password. You will learn that you can create a combination of circles, straight lines, and taps. When ready, press “Choose picture”. Browse your Windows 8.x device and select the picture you want to use for your password and press “Open”. Now you can drag the picture to position it the way you want. When you like how the picture is positioned, press “Use this picture” on the left. If you are not happy with the picture, press “Choose new picture” and select a new one, as shown during the previous step. After you have confirmed that you want to use this picture, you are asked to set up your gestures for the picture password. Draw three gestures on the picture, any combination you wish. Please remember that you can use only three gestures: circles, straight lines, and taps. Once you have drawn those three gestures, you are asked to confirm. Draw the same gestures one more time. If everything goes well, you are informed that you have created your picture password and that you can use it the next time you sign in to Windows. If you don’t confirm the gestures correctly, you will be asked to try again, until you draw the same gestures twice. To close the picture password wizard, press “Finish”. Where Does Windows Store Your Passwords? Are They Safe? All the passwords that you enter in Windows and save for future use are stored in the Credential Manager. This tool is a vault with the usernames and passwords that you use to log on to your computer, to other computers on the network, to apps from the Windows Store, or to websites using Internet Explorer. By storing these credentials, Windows can automatically log you the next time you access the same app, network share, or website. Everything that is stored in the Credential Manager is encrypted for your protection.

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  • An Honest look at SharePoint Web Services

    - by juanlarios
    INTRODUCTION If you are a SharePoint developer you know that there are two basic ways to develop against SharePoint. 1) The object Model 2) Web services. SharePoint object model has the advantage of being quite rich. Anything you can do through the SharePoint UI as an administrator or end user, you can do through the object model. In fact everything that is done through the UI is done through the object model behind the scenes. The major disadvantage to getting at SharePoint this way is that the code needs to run on the server. This means that all web parts, event receivers, features, etc… all of this is code that is deployed to the server. The second way to get to SharePoint is through the built in web services. There are many articles on how to manipulate web services, how to authenticate to them and interact with them. The basic idea is that a remote application or process can contact SharePoint through a web service. Lots has been written about how great these web services are. This article is written to document the limitations, some of the issues and frustrations with working with SharePoint built in web services. Ultimately, for the tasks I was given to , SharePoint built in web services did not suffice. My evaluation of SharePoint built in services was compared against creating my own WCF Services to do what I needed. The current project I'm working on right now involved several "integration points". A remote application, installed on a separate server was to contact SharePoint and perform an task or operation. So I decided to start up Visual Studio and built a DLL and basically have 2 layers of logic. An integration layer and a data layer. A good friend of mine pointed me to SOLID principles and referred me to some videos and tutorials about it. I decided to implement the methodology (although a lot of the principles are common sense and I already incorporated in my coding practices). I was to deliver this dll to the application team and they would simply call the methods exposed by this dll and voila! it would do some task or operation in SharePoint. SOLUTION My integration layer implemented an interface that defined some of the basic integration tasks that I was to put together. My data layer was about the same, it implemented an interface with some of the tasks that I was going to develop. This gave me the opportunity to develop different data layers, ultimately different ways to get at SharePoint if I needed to. This is a classic SOLID principle. In this case it proved to be quite helpful because I wrote one data layer completely implementing SharePoint built in Web Services and another implementing my own WCF Service that I wrote. I should mention there is another layer underneath the data layer. In referencing SharePoint or WCF services in my visual studio project I created a class for every web service call. So for example, if I used List.asx. I created a class called "DocumentRetreival" this class would do the grunt work to connect to the correct URL, It would perform the basic operation of contacting the service and so on. If I used a view.asmx, I implemented a class called "ViewRetrieval" with the same idea as the last class but it would now interact with all he operations in view.asmx. This gave my data layer the ability to perform multiple calls without really worrying about some of the grunt work each class performs. This again, is a classic SOLID principle. So, in order to compare them side by side we can look at both data layers and with is involved in each. Lets take a look at the "Create Project" task or operation. The integration point is described as , "dll is to provide a way to create a project in SharePoint". Projects , in this case are basically document libraries. I am to implement a way in which a remote application can create a document library in SharePoint. Easy enough right? Use the list.asmx Web service in SharePoint. So here we go! Lets take a look at the code. I added the List.asmx web service reference to my project and this is the class that contacts it:  class DocumentRetrieval     {         private ListsSoapClient _service;      d   private bool _impersonation;         public DocumentRetrieval(bool impersonation, string endpt)         {             _service = new ListsSoapClient();             this.SetEndPoint(string.Format("{0}/{1}", endpt, ConfigurationManager.AppSettings["List"]));             _impersonation = impersonation;             if (_impersonation)             {                 _service.ClientCredentials.Windows.ClientCredential.Password = ConfigurationManager.AppSettings["password"];                 _service.ClientCredentials.Windows.ClientCredential.UserName = ConfigurationManager.AppSettings["username"];                 _service.ClientCredentials.Windows.AllowedImpersonationLevel =                     System.Security.Principal.TokenImpersonationLevel.Impersonation;             }     private void SetEndPoint(string p)          {             _service.Endpoint.Address = new EndpointAddress(p);          }          /// <summary>         /// Creates a document library with specific name and templateID         /// </summary>         /// <param name="listName">New list name</param>         /// <param name="templateID">Template ID</param>         /// <returns></returns>         public XmlElement CreateLibrary(string listName, int templateID, ref ExceptionContract exContract)         {             XmlDocument sample = new XmlDocument();             XmlElement viewCol = sample.CreateElement("Empty");             try             {                 _service.Open();                 viewCol = _service.AddList(listName, "", templateID);             }             catch (Exception ex)             {                 exContract = new ExceptionContract("DocumentRetrieval/CreateLibrary", ex.GetType(), "Connection Error", ex.StackTrace, ExceptionContract.ExceptionCode.error);                             }finally             {                 _service.Close();             }                                      return viewCol;         } } There was a lot more in this class (that I am not including) because i was reusing the grunt work and making other operations with LIst.asmx, For example, updating content types, changing or configuring lists or document libraries. One of the first things I noticed about working with the built in services is that you are really at the mercy of what is available to you. Before creating a document library (Project) I wanted to expose a IsProjectExisting method. This way the integration or data layer could recognize if a library already exists. Well there is no service call or method available to do that check. So this is what I wrote:   public bool DocLibExists(string listName, ref ExceptionContract exContract)         {             try             {                 var allLists = _service.GetListCollection();                                return allLists.ChildNodes.OfType<XmlElement>().ToList().Exists(x => x.Attributes["Title"].Value ==listName);             }             catch (Exception ex)             {                 exContract = new ExceptionContract("DocumentRetrieval/GetList/GetListWSCall", ex.GetType(), "Unable to Retrieve List Collection", ex.StackTrace, ExceptionContract.ExceptionCode.error);             }             return false;         } This really just gets an XMLElement with all the lists. It was then up to me to sift through the clutter and noise and see if Document library already existed. This took a little bit of getting used to. Now instead of working with code, you are working with XMLElement response format from web service. I wrote a LINQ query to go through and find if the attribute "Title" existed and had a value of the listname then it would return True, if not False. I didn't particularly like working this way. Dealing with XMLElement responses and then having to manipulate it to get at the exact data I was looking for. Once the check for the DocLibExists, was done, I would either create the document library or send back an error indicating the document library already existed. Now lets examine the code that actually creates the document library. It does what you are really after, it creates a document library. Notice how the template ID is really an integer. Every document library template in SharePoint has an ID associated with it. Document libraries, Image Library, Custom List, Project Tasks, etc… they all he a unique integer associated with it. Well, that's great but the client came back to me and gave me some specifics that each "project" or document library, should have. They specified they had 3 types of projects. Each project would have unique views, about 10 views for each project. Each Project specified unique configurations (auditing, versioning, content types, etc…) So what turned out to be a simple implementation of creating a document library as a repository for a project, turned out to be quite involved.  The first thing I thought of was to create a template for document library. There are other ways you can do this too. Using the web Service call, you could configure views, versioning, even content types, etc… the only catch is, you have to be working quite extensively with CAML. I am not fond of CAML. I can do it and work with it, I just don't like doing it. It is quite touchy and at times it is quite tough to understand where errors were made with CAML statements. Working with Web Services and CAML proved to be quite annoying. The service call would return a generic error message that did not particularly point me to a CAML statement syntax error, or even a CAML error. I was not sure if it was a security , performance or code based issue. It was quite tough to work with. At times it was difficult to work with because of the way SharePoint handles metadata. There are "Names", "Display Name", and "StaticName" fields. It was quite tough to understand at times, which one to use. So it took a lot of trial and error. There are tools that can help with CAML generation. There is also now intellisense for CAML statements in Visual Studio that might help but ultimately I'm not fond of CAML with Web Services.   So I decided on the template. So my plan was to create create a document library, configure it accordingly and then use The Template Builder that comes with the SharePoint SDK. This tool allows you to create site templates, list template etc… It is quite interesting because it does not generate an STP file, it actually generates an xml definition and a feature you can activate and make that template available on a site or site collection. The first issue I experienced with this is that one of the specifications to this template was that the "All Documents" view was to have 2 web parts on it. Well, it turns out that using the template builder , it did not include the web parts as part of the list template definition it generated. It backed up the settings, the views, the content types but not the custom web parts. I still decided to try this even without the web parts on the page. This new template defined a new Document library definition with a unique ID. The problem was that the service call accepts an int but it only has access to the built in library int definitions. Any new ones added or created will not be available to create. So this made it impossible for me to approach the problem this way.     I should also mention that one of the nice features about SharePoint is the ability to create list templates, back them up and then create lists based on that template. It can all be done by end user administrators. These templates are quite unique because they are saved as an STP file and not an xml definition. I also went this route and tried to see if there was another service call where I could create a document library based no given template name. Nope! none.      After some thinking I decide to implement a WCF service to do this creation for me. I was quite certain that the object model would allow me to create document libraries base on a template in which an ID was required and also templates saved as STP files. Now I don't want to bother with posting the code to contact WCF service because it's self explanatory, but I will post the code that I used to create a list with custom template. public ServiceResult CreateProject(string name, string templateName, string projectId)         {             string siteurl = SPContext.Current.Site.Url;             Guid webguid = SPContext.Current.Web.ID;                        using (SPSite site = new SPSite(siteurl))             {                 using (SPWeb rootweb = site.RootWeb)                 {                     SPListTemplateCollection temps = site.GetCustomListTemplates(rootweb);                     ProcessWeb(siteurl, webguid, web => Act_CreateProject(web, name, templateName, projectId, temps));                 }//SpWeb             }//SPSite              return _globalResult;                   }         private void Act_CreateProject(SPWeb targetsite, string name, string templateName, string projectId, SPListTemplateCollection temps) {                         var temp = temps.Cast<SPListTemplate>().FirstOrDefault(x => x.Name.Equals(templateName));             if (temp != null)             {                             try                 {                                         Guid listGuid = targetsite.Lists.Add(name, "", temp);                     SPList newList = targetsite.Lists[listGuid];                     _globalResult = new ServiceResult(true, "Success", "Success");                 }                 catch (Exception ex)                 {                     _globalResult = new ServiceResult(false, (string.IsNullOrEmpty(ex.Message) ? "None" : ex.Message + " " + templateName), ex.StackTrace.ToString());                 }                                       }        private void ProcessWeb(string siteurl, Guid webguid, Action<SPWeb> action) {                        using (SPSite sitecollection = new SPSite(siteurl)) {                 using (SPWeb web = sitecollection.AllWebs[webguid]) {                     action(web);                 }                     }                  } This code is actually some of the code I implemented for the service. there was a lot more I did on Project Creation which I will cover in my next blog post. I implemented an ACTION method to process the web. This allowed me to properly dispose the SPWEb and SPSite objects and not rewrite this code over and over again. So I implemented a WCF service to create projects for me, this allowed me to do a lot more than just create a document library with a template, it now gave me the flexibility to do just about anything the client wanted at project creation. Once this was implemented , the client came back to me and said, "we reference all our projects with ID's in our application. we want SharePoint to do the same". This has been something I have been doing for a little while now but I do hope that SharePoint 2010 can have more of an answer to this and address it properly. I have been adding metadata to SPWebs through property bag. I believe I have blogged about it before. This time it required metadata added to a document library. No problem!!! I also mentioned these web parts that were to go on the "All Documents" View. I took the opportunity to configure them to the appropriate settings. There were two settings that needed to be set on these web parts. One of them was a Project ID configured in the webpart properties. The following code enhances and replaces the "Act_CreateProject " method above:  private void Act_CreateProject(SPWeb targetsite, string name, string templateName, string projectId, SPListTemplateCollection temps) {                         var temp = temps.Cast<SPListTemplate>().FirstOrDefault(x => x.Name.Equals(templateName));             if (temp != null)             {                 SPLimitedWebPartManager wpmgr = null;                               try                 {                                         Guid listGuid = targetsite.Lists.Add(name, "", temp);                     SPList newList = targetsite.Lists[listGuid];                     SPFolder rootFolder = newList.RootFolder;                     rootFolder.Properties.Add(KEY, projectId);                     rootFolder.Update();                     if (rootFolder.ParentWeb != targetsite)                         rootFolder.ParentWeb.Dispose();                     if (!templateName.Contains("Natural"))                     {                         SPView alldocumentsview = newList.Views.Cast<SPView>().FirstOrDefault(x => x.Title.Equals(ALLDOCUMENTS));                         SPFile alldocfile = targetsite.GetFile(alldocumentsview.ServerRelativeUrl);                         wpmgr = alldocfile.GetLimitedWebPartManager(PersonalizationScope.Shared);                         ConfigureWebPart(wpmgr, projectId, CUSTOMWPNAME);                                              alldocfile.Update();                     }                                        if (newList.ParentWeb != targetsite)                         newList.ParentWeb.Dispose();                     _globalResult = new ServiceResult(true, "Success", "Success");                 }                 catch (Exception ex)                 {                     _globalResult = new ServiceResult(false, (string.IsNullOrEmpty(ex.Message) ? "None" : ex.Message + " " + templateName), ex.StackTrace.ToString());                 }                 finally                 {                     if (wpmgr != null)                     {                         wpmgr.Web.Dispose();                         wpmgr.Dispose();                     }                 }             }                         }       private void ConfigureWebPart(SPLimitedWebPartManager mgr, string prjId, string webpartname)         {             var wp = mgr.WebParts.Cast<System.Web.UI.WebControls.WebParts.WebPart>().FirstOrDefault(x => x.DisplayTitle.Equals(webpartname));             if (wp != null)             {                           (wp as ListRelationshipWebPart.ListRelationshipWebPart).ProjectID = prjId;                 mgr.SaveChanges(wp);             }         }   This Shows you how I was able to set metadata on the document library. It has to be added to the RootFolder of the document library, Unfortunately, the SPList does not have a Property bag that I can add a key\value pair to. It has to be done on the root folder. Now everything in the integration will reference projects by ID's and will not care about names. My, "DocLibExists" will now need to be changed because a web service is not set up to look at property bags.  I had to write another method on the Service to do the equivalent but with ID's instead of names.  The second thing you will notice about the code is the use of the Webpartmanager. I have seen several examples online, and also read a lot about memory leaks, The above code does not produce memory leaks. The web part manager creates an SPWeb, so just dispose it like I did. CONCLUSION This is a long long post so I will stop here for now, I will continue with more comparisons and limitations in my next post. My conclusion for this example is that Web Services will do the trick if you can suffer through CAML and if you are doing some simple operations. For Everything else, there's WCF! **** fireI apologize for the disorganization of this post, I was on a bus on a 12 hour trip to IOWA while I wrote it, I was half asleep and half awake, hopefully it makes enough sense to someone.

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  • Metro: Query Selectors

    - by Stephen.Walther
    The goal of this blog entry is to explain how to perform queries using selectors when using the WinJS library. In particular, you learn how to use the WinJS.Utilities.query() method and the QueryCollection class to retrieve and modify the elements of an HTML document. Introduction to Selectors When you are building a Web application, you need some way of easily retrieving elements from an HTML document. For example, you might want to retrieve all of the input elements which have a certain class. Or, you might want to retrieve the one and only element with an id of favoriteColor. The standard way of retrieving elements from an HTML document is by using a selector. Anyone who has ever created a Cascading Style Sheet has already used selectors. You use selectors in Cascading Style Sheets to apply formatting rules to elements in a document. For example, the following Cascading Style Sheet rule changes the background color of every INPUT element with a class of .required in a document to the color red: input.red { background-color: red } The “input.red” part is the selector which matches all INPUT elements with a class of red. The W3C standard for selectors (technically, their recommendation) is entitled “Selectors Level 3” and the standard is located here: http://www.w3.org/TR/css3-selectors/ Selectors are not only useful for adding formatting to the elements of a document. Selectors are also useful when you need to apply behavior to the elements of a document. For example, you might want to select a particular BUTTON element with a selector and add a click handler to the element so that something happens whenever you click the button. Selectors are not specific to Cascading Style Sheets. You can use selectors in your JavaScript code to retrieve elements from an HTML document. jQuery is famous for its support for selectors. Using jQuery, you can use a selector to retrieve matching elements from a document and modify the elements. The WinJS library enables you to perform the same types of queries as jQuery using the W3C selector syntax. Performing Queries with the WinJS.Utilities.query() Method When using the WinJS library, you perform a query using a selector by using the WinJS.Utilities.query() method.  The following HTML document contains a BUTTON and a DIV element: <!DOCTYPE html> <html> <head> <meta charset="utf-8"> <title>Application1</title> <!-- WinJS references --> <link href="//Microsoft.WinJS.0.6/css/ui-dark.css" rel="stylesheet"> <script src="//Microsoft.WinJS.0.6/js/base.js"></script> <script src="//Microsoft.WinJS.0.6/js/ui.js"></script> <!-- Application1 references --> <link href="/css/default.css" rel="stylesheet"> <script src="/js/default.js"></script> </head> <body> <button>Click Me!</button> <div style="display:none"> <h1>Secret Message</h1> </div> </body> </html> The document contains a reference to the following JavaScript file named \js\default.js: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { WinJS.Utilities.query("button").listen("click", function () { WinJS.Utilities.query("div").clearStyle("display"); }); } }; app.start(); })(); The default.js script uses the WinJS.Utilities.query() method to retrieve all of the BUTTON elements in the page. The listen() method is used to wire an event handler to the BUTTON click event. When you click the BUTTON, the secret message contained in the hidden DIV element is displayed. The clearStyle() method is used to remove the display:none style attribute from the DIV element. Under the covers, the WinJS.Utilities.query() method uses the standard querySelectorAll() method. This means that you can use any selector which is compatible with the querySelectorAll() method when using the WinJS.Utilities.query() method. The querySelectorAll() method is defined in the W3C Selectors API Level 1 standard located here: http://www.w3.org/TR/selectors-api/ Unlike the querySelectorAll() method, the WinJS.Utilities.query() method returns a QueryCollection. We talk about the methods of the QueryCollection class below. Retrieving a Single Element with the WinJS.Utilities.id() Method If you want to retrieve a single element from a document, instead of matching a set of elements, then you can use the WinJS.Utilities.id() method. For example, the following line of code changes the background color of an element to the color red: WinJS.Utilities.id("message").setStyle("background-color", "red"); The statement above matches the one and only element with an Id of message. For example, the statement matches the following DIV element: <div id="message">Hello!</div> Notice that you do not use a hash when matching a single element with the WinJS.Utilities.id() method. You would need to use a hash when using the WinJS.Utilities.query() method to do the same thing like this: WinJS.Utilities.query("#message").setStyle("background-color", "red"); Under the covers, the WinJS.Utilities.id() method calls the standard document.getElementById() method. The WinJS.Utilities.id() method returns the result as a QueryCollection. If no element matches the identifier passed to WinJS.Utilities.id() then you do not get an error. Instead, you get a QueryCollection with no elements (length=0). Using the WinJS.Utilities.children() method The WinJS.Utilities.children() method enables you to retrieve a QueryCollection which contains all of the children of a DOM element. For example, imagine that you have a DIV element which contains children DIV elements like this: <div id="discussContainer"> <div>Message 1</div> <div>Message 2</div> <div>Message 3</div> </div> You can use the following code to add borders around all of the child DIV element and not the container DIV element: var discussContainer = WinJS.Utilities.id("discussContainer").get(0); WinJS.Utilities.children(discussContainer).setStyle("border", "2px dashed red");   It is important to understand that the WinJS.Utilities.children() method only works with a DOM element and not a QueryCollection. Notice that the get() method is used to retrieve the DOM element which represents the discussContainer. Working with the QueryCollection Class Both the WinJS.Utilities.query() method and the WinJS.Utilities.id() method return an instance of the QueryCollection class. The QueryCollection class derives from the base JavaScript Array class and adds several useful methods for working with HTML elements: addClass(name) – Adds a class to every element in the QueryCollection. clearStyle(name) – Removes a style from every element in the QueryCollection. conrols(ctor, options) – Enables you to create controls. get(index) – Retrieves the element from the QueryCollection at the specified index. getAttribute(name) – Retrieves the value of an attribute for the first element in the QueryCollection. hasClass(name) – Returns true if the first element in the QueryCollection has a certain class. include(items) – Includes a collection of items in the QueryCollection. listen(eventType, listener, capture) – Adds an event listener to every element in the QueryCollection. query(query) – Performs an additional query on the QueryCollection and returns a new QueryCollection. removeClass(name) – Removes a class from the every element in the QueryCollection. removeEventListener(eventType, listener, capture) – Removes an event listener from every element in the QueryCollection. setAttribute(name, value) – Adds an attribute to every element in the QueryCollection. setStyle(name, value) – Adds a style attribute to every element in the QueryCollection. template(templateElement, data, renderDonePromiseContract) – Renders a template using the supplied data.  toggleClass(name) – Toggles the specified class for every element in the QueryCollection. Because the QueryCollection class derives from the base Array class, it also contains all of the standard Array methods like forEach() and slice(). Summary In this blog post, I’ve described how you can perform queries using selectors within a Windows Metro Style application written with JavaScript. You learned how to return an instance of the QueryCollection class by using the WinJS.Utilities.query(), WinJS.Utilities.id(), and WinJS.Utilities.children() methods. You also learned about the methods of the QueryCollection class.

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  • Option Trading: Getting the most out of the event session options

    - by extended_events
    You can control different aspects of how an event session behaves by setting the event session options as part of the CREATE EVENT SESSION DDL. The default settings for the event session options are designed to handle most of the common event collection situations so I generally recommend that you just use the defaults. Like everything in the real world though, there are going to be a handful of “special cases” that require something different. This post focuses on identifying the special cases and the correct use of the options to accommodate those cases. There is a reason it’s called Default The default session options specify a total event buffer size of 4 MB with a 30 second latency. Translating this into human terms; this means that our default behavior is that the system will start processing events from the event buffer when we reach about 1.3 MB of events or after 30 seconds, which ever comes first. Aside: What’s up with the 1.3 MB, I thought you said the buffer was 4 MB?The Extended Events engine takes the total buffer size specified by MAX_MEMORY (4MB by default) and divides it into 3 equally sized buffers. This is done so that a session can be publishing events to one buffer while other buffers are being processed. There are always at least three buffers; how to get more than three is covered later. Using this configuration, the Extended Events engine can “keep up” with most event sessions on standard workloads. Why is this? The fact is that most events are small, really small; on the order of a couple hundred bytes. Even when you start considering events that carry dynamically sized data (eg. binary, text, etc.) or adding actions that collect additional data, the total size of the event is still likely to be pretty small. This means that each buffer can likely hold thousands of events before it has to be processed. When the event buffers are finally processed there is an economy of scale achieved since most targets support bulk processing of the events so they are processed at the buffer level rather than the individual event level. When all this is working together it’s more likely that a full buffer will be processed and put back into the ready queue before the remaining buffers (remember, there are at least three) are full. I know what you’re going to say: “My server is exceptional! My workload is so massive it defies categorization!” OK, maybe you weren’t going to say that exactly, but you were probably thinking it. The point is that there are situations that won’t be covered by the Default, but that’s a good place to start and this post assumes you’ve started there so that you have something to look at in order to determine if you do have a special case that needs different settings. So let’s get to the special cases… What event just fired?! How about now?! Now?! If you believe the commercial adage from Heinz Ketchup (Heinz Slow Good Ketchup ad on You Tube), some things are worth the wait. This is not a belief held by most DBAs, particularly DBAs who are looking for an answer to a troubleshooting question fast. If you’re one of these anxious DBAs, or maybe just a Program Manager doing a demo, then 30 seconds might be longer than you’re comfortable waiting. If you find yourself in this situation then consider changing the MAX_DISPATCH_LATENCY option for your event session. This option will force the event buffers to be processed based on your time schedule. This option only makes sense for the asynchronous targets since those are the ones where we allow events to build up in the event buffer – if you’re using one of the synchronous targets this option isn’t relevant. Avoid forgotten events by increasing your memory Have you ever had one of those days where you keep forgetting things? That can happen in Extended Events too; we call it dropped events. In order to optimizes for server performance and help ensure that the Extended Events doesn’t block the server if to drop events that can’t be published to a buffer because the buffer is full. You can determine if events are being dropped from a session by querying the dm_xe_sessions DMV and looking at the dropped_event_count field. Aside: Should you care if you’re dropping events?Maybe not – think about why you’re collecting data in the first place and whether you’re really going to miss a few dropped events. For example, if you’re collecting query duration stats over thousands of executions of a query it won’t make a huge difference to miss a couple executions. Use your best judgment. If you find that your session is dropping events it means that the event buffer is not large enough to handle the volume of events that are being published. There are two ways to address this problem. First, you could collect fewer events – examine you session to see if you are over collecting. Do you need all the actions you’ve specified? Could you apply a predicate to be more specific about when you fire the event? Assuming the session is defined correctly, the next option is to change the MAX_MEMORY option to a larger number. Picking the right event buffer size might take some trial and error, but a good place to start is with the number of dropped events compared to the number you’ve collected. Aside: There are three different behaviors for dropping events that you specify using the EVENT_RETENTION_MODE option. The default is to allow single event loss and you should stick with this setting since it is the best choice for keeping the impact on server performance low.You’ll be tempted to use the setting to not lose any events (NO_EVENT_LOSS) – resist this urge since it can result in blocking on the server. If you’re worried that you’re losing events you should be increasing your event buffer memory as described in this section. Some events are too big to fail A less common reason for dropping an event is when an event is so large that it can’t fit into the event buffer. Even though most events are going to be small, you might find a condition that occasionally generates a very large event. You can determine if your session is dropping large events by looking at the dm_xe_sessions DMV once again, this time check the largest_event_dropped_size. If this value is larger than the size of your event buffer [remember, the size of your event buffer, by default, is max_memory / 3] then you need a large event buffer. To specify a large event buffer you set the MAX_EVENT_SIZE option to a value large enough to fit the largest event dropped based on data from the DMV. When you set this option the Extended Events engine will create two buffers of this size to accommodate these large events. As an added bonus (no extra charge) the large event buffer will also be used to store normal events in the cases where the normal event buffers are all full and waiting to be processed. (Note: This is just a side-effect, not the intended use. If you’re dropping many normal events then you should increase your normal event buffer size.) Partitioning: moving your events to a sub-division Earlier I alluded to the fact that you can configure your event session to use more than the standard three event buffers – this is called partitioning and is controlled by the MEMORY_PARTITION_MODE option. The result of setting this option is fairly easy to explain, but knowing when to use it is a bit more art than science. First the science… You can configure partitioning in three ways: None, Per NUMA Node & Per CPU. This specifies the location where sets of event buffers are created with fairly obvious implication. There are rules we follow for sub-dividing the total memory (specified by MAX_MEMORY) between all the event buffers that are specific to the mode used: None: 3 buffers (fixed)Node: 3 * number_of_nodesCPU: 2.5 * number_of_cpus Here are some examples of what this means for different Node/CPU counts: Configuration None Node CPU 2 CPUs, 1 Node 3 buffers 3 buffers 5 buffers 6 CPUs, 2 Node 3 buffers 6 buffers 15 buffers 40 CPUs, 5 Nodes 3 buffers 15 buffers 100 buffers   Aside: Buffer size on multi-processor computersAs the number of Nodes or CPUs increases, the size of the event buffer gets smaller because the total memory is sub-divided into more pieces. The defaults will hold up to this for a while since each buffer set is holding events only from the Node or CPU that it is associated with, but at some point the buffers will get too small and you’ll either see events being dropped or you’ll get an error when you create your session because you’re below the minimum buffer size. Increase the MAX_MEMORY setting to an appropriate number for the configuration. The most likely reason to start partitioning is going to be related to performance. If you notice that running an event session is impacting the performance of your server beyond a reasonably expected level [Yes, there is a reasonably expected level of work required to collect events.] then partitioning might be an answer. Before you partition you might want to check a few other things: Is your event retention set to NO_EVENT_LOSS and causing blocking? (I told you not to do this.) Consider changing your event loss mode or increasing memory. Are you over collecting and causing more work than necessary? Consider adding predicates to events or removing unnecessary events and actions from your session. Are you writing the file target to the same slow disk that you use for TempDB and your other high activity databases? <kidding> <not really> It’s always worth considering the end to end picture – if you’re writing events to a file you can be impacted by I/O, network; all the usual stuff. Assuming you’ve ruled out the obvious (and not so obvious) issues, there are performance conditions that will be addressed by partitioning. For example, it’s possible to have a successful event session (eg. no dropped events) but still see a performance impact because you have many CPUs all attempting to write to the same free buffer and having to wait in line to finish their work. This is a case where partitioning would relieve the contention between the different CPUs and likely reduce the performance impact cause by the event session. There is no DMV you can check to find these conditions – sorry – that’s where the art comes in. This is  largely a matter of experimentation. On the bright side you probably won’t need to to worry about this level of detail all that often. The performance impact of Extended Events is significantly lower than what you may be used to with SQL Trace. You will likely only care about the impact if you are trying to set up a long running event session that will be part of your everyday workload – sessions used for short term troubleshooting will likely fall into the “reasonably expected impact” category. Hey buddy – I think you forgot something OK, there are two options I didn’t cover: STARTUP_STATE & TRACK_CAUSALITY. If you want your event sessions to start automatically when the server starts, set the STARTUP_STATE option to ON. (Now there is only one option I didn’t cover.) I’m going to leave causality for another post since it’s not really related to session behavior, it’s more about event analysis. - Mike Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • Your Day-by-Day Guide to Agile PLM at Oracle OpenWorld 2012

    - by Kerrie Foy
    This year’s Oracle OpenWorld conference is nearly here, and we’re all excited about what we have planned! With five days of activities and customer presenters from market leaders and top innovators like The Coca-Cola Company, Starbucks, JDSU, Facebook, GlobalFoundries, and more, this is an event you don't want to miss. I've compiled this day-by-day guide to help anyone keep track of all the “Product Lifecycle Management and Product Value Chain” sessions and activities at OpenWorld 2012, September 30 – October 4 in San Francisco, California.  Monday, October 1 There are great networking activities on Sunday September 30, but PLM specific sessions start after general conference keynotes on Monday, October 1 at 10:45 a.m. at the InterContinental Hotel in room Telegraph Hill. In fact, most of our sessions this year will be held in this room, which is still close to the conference keynotes in Moscone, but just far enough away to allow some focused networking and discussions.   This first session, 10:45 – 11:45 a.m. is a joint session with the Agile and AutoVue teams, entitled “Streamline PLM Design-to-Manufacturing Processes with AutoVue Visualization Soltuions” featuring presenters from Oracle as well as joint AutoVue and Agile PLM customer GlobalFoundries. In the following 12:15 – 1:15 p.m. slot, there are two sessions to choose from, so if you have a team of representatives attending OpenWorld, you may consider splitting up to catch both of these: a) Our General Session will be held in the InterContinental Hotel Ballroom C, which will cover our complete enterprise PLM strategy, product updates, and roadmaps. It’s our pleasure to feature a customer keynote presentation from Chris Bedi, CIO, and Rajeev Sethi, Director IT Business Engagement, of JDSU. b) A focused session on integrating PLM with Engineering and Supply Chain Systems will be held on the second floor of Moscone West (next to the InterContinental) in room 2022. Join to discover how these types of integrations help companies manage common and integrated design information across all MCAD, ECAD, and software components. After a lunch break and perhaps a visit to the Demogrounds in Moscone West, select from two product roadmap sessions in the next time slot (3:15 – 4:15 p.m.): an Agile 9.3.x session located in the InterContinental’s Ballroom C, and an Agile PLM for Process session located back in the InterContinental’s Telegraph Room. Both sessions will have strong content around each product line’s latest releases, vision, and customer examples. We are very pleased to feature Daniel Soosai of Facebook in the A9 session and Vinnie D’Agostino of The Coca-Cola Company in the PLM for Process session. Afterwards, hang in there for one last session of the day from 4:45 – 5:45 p.m.; it’s an insightful discussion on leveraging Agile PLM as the Foundation for Enterprise Quality Management, and it’s sure to be one of the best. In the Telegraph Room, this session will feature Oracle experts, partner co-presenter David Bartlett from CPG Solutions, and customer co-presenter Thomas Crowe, CIO of PL Developments. Hear their experience around implementing collaborative, integrated solutions to ensure effective knowledge transfer throughout an organization, and how to perform analysis in real time to resolve product quality issues swiftly and efficiently. On Monday evening there will be plenty of industry, product, and partner dinners, so take advantage of all the networking opportunities and catch some great tunes at the 5 day Oracle OpenWorld Music Festival! Tuesday, October 2 Tuesday starts early with a special PLM Networking Brunch, sponsored by several partners, from 8:30 a.m. – 10:30 a.m. at the B Restaurant that sits atop Yerba Buena Gardens. You’ll have the unique opportunity to meet with like-minded industry peers and a PLM partner to discuss a topic of your choosing while enjoying a delicious meal. Registration is required, so to inquire about attending this brunch, please email Terri.Hiskey-AT-oracle.com. After wrapping up your conversations over brunch, head over to the Marriott Marquis in the Nob Hill CD room for a chance to experience the Oracle Product Lifecycle Analytics solution in a Hands-On Lab, open from 10:15 a.m. – 12:45 p.m. Experts will be there to answer your questions. Back in the InterContinental Hotel’s Telegraph room, the session on “Ideation and Requirements Management: Capturing the Voice of the Customer” begins at 11:45 a.m. – 12:45 p.m. This may be the session for you if you’re struggling with challenges like too many repositories of customer needs, requests, and ideas; limited visibility into which ideas are being advanced by customers and field resources; or if you’re unable to leverage internal expertise to expose effort and potential risks. This session will discuss how Agile PLM can help you overcome ideation challenges to deliver the right products to their targeted markets and fulfill customer desires. Next, from 1:15 – 2:15 p.m. join us for a session on Managing Profitable Innovation with Oracle Product Lifecycle Analytics. If you missed the Hands-on Lab, have more questions, or simply want to be inspired by the product’s forward-thinking vision and capabilities, this is a great opportunity to meet the progressive-minded executives behind the application. After this session, it may be a good opportunity to swing by the Demogrounds in Moscone West and visit the Agile PLM demos at exhibit booths #81 for Agile PLM for Discrete Manufacturing, #70 for Agile PLM for Process, and #82 for AutoVue and Agile PLM Enterprise Visualization. Check out the related Supply Chain Management booths close by if you’re interested - here's the map. There’s always lots to see and do around the exhibit area. But don’t forget the last session of the day from 5:00 p.m. – 6:00 p.m. in Telegraph Hill on Managing Product Innovation and Compliance in Life Science Companies, a “must-see” if you’re in this industry. Launching innovative products quickly is already a high-stakes challenge, but companies in the life sciences industry face uniquely severe consequences when new products don’t perform or comply as required. In recent years, more and more regulations have become mandatory, and new ones, such as REACH, are currently going into effect for several companies. Customer presenters from pharmaceutical leader Eli Lilly will share how they’ve leveraged Agile PLM to deliver high-quality, innovative products in a fast-paced, heavily regulated market environment. Tuesday evening unwind at the Supply Chain Management Reception from 6:00 – 8:00 p.m. at the premier boutique Roe Nightclub and Lounge, which is located about three blocks down on Howard Street (on the other side of Moscone from the InterContinental Hotel). Registration is required. Click here for the details.   Wednesday, October 3 We have another full line-up on Wednesday, so be ready for an action-packed day. We start with a session at 10:15 – 11:15 a.m. in the Telegraph Room where we have a session on “PLM for Consumer Products: Building an Engine for Quality and Innovation” with featured presenters from Starbucks and partner Kalypso. This is a rare opportunity to learn directly from Starbucks how they instill quality and innovation throughout their organization, products, and processes, leveraging PLM disciplines with strong support from their partner.  If you’re not in the consumer products industry, we recommend attending another session at 10:15 – 11:15 a.m. in Moscone West room 3005: “Eco-Enterprise Innovation Awards and the Business Case for Sustainability” featuring Jeff Henley, Oracle’s Chairman of the Board and Jon Chorley, Chief Sustainability Officer. Oracle will honor select customers with Oracle’s Eco-Enterprise Innovation award, which recognizes customers and their respective partners who rely on Oracle products to support their green business practices to reduce their environmental impact while improving business efficiencies and reducing costs. The awards presentation is followed by a panel discussion with customers and Oracle executives, who describe how these award-winning organizations are embracing environmental initiatives as a central part of their business strategy and how information technology plays a pivotal role. Next at 11:45 a.m. – 12:45 p.m. in Telegraph Hill attend our session devoted to exploring Product Lifecycle Management’s role in Software Lifecycle Management. This is a thought leadership session with Oracle experts in the field on the importance of change management, and we’ll discuss how Oracle has for years leveraged Agile PLM to develop Agile PLM. If software lifecycle management doesn’t apply to your business or you’d rather engage in some lively one-on-one discussions, we also have a “Supply Chain Meet the Experts” session in Moscone West Room 2001A. Product experts, thought leaders and executives will be on hand to discuss your questions/topics, so come prepared. This session tends to fill up fast so try to get in early. At 1:15 – 2:15 p.m. join us back in Telegraph Hill for a session focused on leveraging the Agile Product Portfolio Management application as the Product Development Master Schedule to improve efficiencies, optimize resources, and gain visibility across projects enterprise-wide to improve portfolio profitability. Customer presenters from Broadcom will explain how they’ve leveraged the product to enable a master schedule with enterprise-level, phase-gate program and project collaboration and resource optimization. Again in Telegraph Hill from 3:30 – 4:30 p.m. we have an interesting session with leading semiconductor customer LSI and partner Kalypso on how LSI leveraged Agile PLM to advance from homegrown applications to complete Product Value Chain Management. That type of transition can be challenging, and LSI details how they were able to achieve their goals and the value they gained along the journey – a fascinating account for any company interested in leveraging best practices to innovate their business processes and even end products. Lastly, we’ll wrap up in Telegraph Hill from 5:00 – 6:00 p.m. with a session on “Ensuring New Product Success by Achieving Excellence in New Product Introduction.” This is a cross-industry session, guaranteed to deliver insight in the often elusive practice of creating winning products, and we’re very excited about. According to IDC Manufacturing Insights analyst Joe Barkai, “Product Failures are not necessarily a result of bad ideas…they are a result of suboptimal decisions.” We’ll show you how to wire your business processes to enhance decision-making and maximize product potential. Now, quickly hit your hotel room to freshen up and then catch one of the many complimentary shuttles to the much-anticipated Oracle Customer Appreciation Event on Treasure Island. We have a very exciting show planned – check out what’s in store here. Thursday, October 4 PLM has a light schedule on Thursday this year with just one session, but this again is one of our best sessions on managing the Product Value Chain: at 11:15 a.m – 12:15 p.m.in Telegraph Hill, it’s a customer and partner driven session with Sonoco Products and Deloitte telling their story about how to achieve integrated change control by interfacing Agile PLM with Oracle E-Business Suite. Sonoco Products, a global manufacturer of consumer and industrial packaging materials, with its systems integrator, Deloitte, is doing this by implementing prebuilt integration (Oracle Design-to-Release Integration Pack for Agile Product Lifecycle Management for Process and Oracle Process) to integrate Agile with Oracle Product Hub/Oracle Product Information Management and Oracle E-Business Suite. This session presents a case study of how Sonoco is leveraging this solution to improve data quality and build a framework for stronger master data governance. Even though that ends our PLM line-up at OpenWorld, there will still be many sessions and activities at the conference, so visit the Oracle OpenWorld website to review agendas and build your schedule. And of course, download and bring this guide and the latest version of the Agile PLM Focus-On Document (available soon!). San Francisco is a wonderful city to explore, and we’re glad you’re considering joining the Agile PLM team at Oracle OpenWorld!  I hope to see you there! Follow me before the conference and on site for real-time updates about #OOW12 on Twitter @Kerrie_Foy or @AgilePLM.

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  • exporting bind and keyframe bone poses from blender to use in OpenGL

    - by SaldaVonSchwartz
    I'm having a hard time trying to understand how exactly Blender's concept of bone transforms maps to the usual math of skinning (which I'm implementing in an OpenGL-based engine of sorts). Or I'm missing out something in the math.. It's gonna be long, but here's as much background as I can think of. First, a few notes and assumptions: I'm using column-major order and multiply from right to left. So for instance, vertex v transformed by matrix A and then further transformed by matrix B would be: v' = BAv. This also means whenever I export a matrix from blender through python, I export it (in text format) in 4 lines, each representing a column. This is so I can then I can read them back into my engine like this: if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[0], &skeleton.joints[currentJointIndex].inverseBindTransform.m[1], &skeleton.joints[currentJointIndex].inverseBindTransform.m[2], &skeleton.joints[currentJointIndex].inverseBindTransform.m[3])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[4], &skeleton.joints[currentJointIndex].inverseBindTransform.m[5], &skeleton.joints[currentJointIndex].inverseBindTransform.m[6], &skeleton.joints[currentJointIndex].inverseBindTransform.m[7])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[8], &skeleton.joints[currentJointIndex].inverseBindTransform.m[9], &skeleton.joints[currentJointIndex].inverseBindTransform.m[10], &skeleton.joints[currentJointIndex].inverseBindTransform.m[11])) { if (fscanf(fileHandle, "%f %f %f %f", &skeleton.joints[currentJointIndex].inverseBindTransform.m[12], &skeleton.joints[currentJointIndex].inverseBindTransform.m[13], &skeleton.joints[currentJointIndex].inverseBindTransform.m[14], &skeleton.joints[currentJointIndex].inverseBindTransform.m[15])) { I'm simplifying the code I show because otherwise it would make things unnecessarily harder (in the context of my question) to explain / follow. Please refrain from making remarks related to optimizations. This is not final code. Having said that, if I understand correctly, the basic idea of skinning/animation is: I have a a mesh made up of vertices I have the mesh model-world transform W I have my joints, which are really just transforms from each joint's space to its parent's space. I'll call these transforms Bj meaning matrix which takes from joint j's bind pose to joint j-1's bind pose. For each of these, I actually import their inverse to the engine, Bj^-1. I have keyframes each containing a set of current poses Cj for each joint J. These are initially imported to my engine in TQS format but after (S)LERPING them I compose them into Cj matrices which are equivalent to the Bjs (not the Bj^-1 ones) only that for the current spacial configurations of each joint at that frame. Given the above, the "skeletal animation algorithm is" On each frame: check how much time has elpased and compute the resulting current time in the animation, from 0 meaning frame 0 to 1, meaning the end of the animation. (Oh and I'm looping forever so the time is mod(total duration)) for each joint: 1 -calculate its world inverse bind pose, that is Bj_w^-1 = Bj^-1 Bj-1^-1 ... B0^-1 2 -use the current animation time to LERP the componets of the TQS and come up with an interpolated current pose matrix Cj which should transform from the joints current configuration space to world space. Similar to what I did to get the world version of the inverse bind poses, I come up with the joint's world current pose, Cj_w = C0 C1 ... Cj 3 -now that I have world versions of Bj and Cj, I store this joint's world- skinning matrix K_wj = Cj_w Bj_w^-1. The above is roughly implemented like so: - (void)update:(NSTimeInterval)elapsedTime { static double time = 0; time = fmod((time + elapsedTime),1.); uint16_t LERPKeyframeNumber = 60 * time; uint16_t lkeyframeNumber = 0; uint16_t lkeyframeIndex = 0; uint16_t rkeyframeNumber = 0; uint16_t rkeyframeIndex = 0; for (int i = 0; i < aClip.keyframesCount; i++) { uint16_t keyframeNumber = aClip.keyframes[i].number; if (keyframeNumber <= LERPKeyframeNumber) { lkeyframeIndex = i; lkeyframeNumber = keyframeNumber; } else { rkeyframeIndex = i; rkeyframeNumber = keyframeNumber; break; } } double lTime = lkeyframeNumber / 60.; double rTime = rkeyframeNumber / 60.; double blendFactor = (time - lTime) / (rTime - lTime); GLKMatrix4 bindPosePalette[aSkeleton.jointsCount]; GLKMatrix4 currentPosePalette[aSkeleton.jointsCount]; for (int i = 0; i < aSkeleton.jointsCount; i++) { F3DETQSType& lPose = aClip.keyframes[lkeyframeIndex].skeletonPose.jointPoses[i]; F3DETQSType& rPose = aClip.keyframes[rkeyframeIndex].skeletonPose.jointPoses[i]; GLKVector3 LERPTranslation = GLKVector3Lerp(lPose.t, rPose.t, blendFactor); GLKQuaternion SLERPRotation = GLKQuaternionSlerp(lPose.q, rPose.q, blendFactor); GLKVector3 LERPScaling = GLKVector3Lerp(lPose.s, rPose.s, blendFactor); GLKMatrix4 currentTransform = GLKMatrix4MakeWithQuaternion(SLERPRotation); currentTransform = GLKMatrix4Multiply(currentTransform, GLKMatrix4MakeTranslation(LERPTranslation.x, LERPTranslation.y, LERPTranslation.z)); currentTransform = GLKMatrix4Multiply(currentTransform, GLKMatrix4MakeScale(LERPScaling.x, LERPScaling.y, LERPScaling.z)); if (aSkeleton.joints[i].parentIndex == -1) { bindPosePalette[i] = aSkeleton.joints[i].inverseBindTransform; currentPosePalette[i] = currentTransform; } else { bindPosePalette[i] = GLKMatrix4Multiply(aSkeleton.joints[i].inverseBindTransform, bindPosePalette[aSkeleton.joints[i].parentIndex]); currentPosePalette[i] = GLKMatrix4Multiply(currentPosePalette[aSkeleton.joints[i].parentIndex], currentTransform); } aSkeleton.skinningPalette[i] = GLKMatrix4Multiply(currentPosePalette[i], bindPosePalette[i]); } } At this point, I should have my skinning palette. So on each frame in my vertex shader, I do: uniform mat4 modelMatrix; uniform mat4 projectionMatrix; uniform mat3 normalMatrix; uniform mat4 skinningPalette[6]; attribute vec4 position; attribute vec3 normal; attribute vec2 tCoordinates; attribute vec4 jointsWeights; attribute vec4 jointsIndices; varying highp vec2 tCoordinatesVarying; varying highp float lIntensity; void main() { vec3 eyeNormal = normalize(normalMatrix * normal); vec3 lightPosition = vec3(0., 0., 2.); lIntensity = max(0.0, dot(eyeNormal, normalize(lightPosition))); tCoordinatesVarying = tCoordinates; vec4 skinnedVertexPosition = vec4(0.); for (int i = 0; i < 4; i++) { skinnedVertexPosition += jointsWeights[i] * skinningPalette[int(jointsIndices[i])] * position; } gl_Position = projectionMatrix * modelMatrix * skinnedVertexPosition; } The result: The mesh parts that are supposed to animate do animate and follow the expected motion, however, the rotations are messed up in terms of orientations. That is, the mesh is not translated somewhere else or scaled in any way, but the orientations of rotations seem to be off. So a few observations: In the above shader notice I actually did not multiply the vertices by the mesh modelMatrix (the one which would take them to model or world or global space, whichever you prefer, since there is no parent to the mesh itself other than "the world") until after skinning. This is contrary to what I implied in the theory: if my skinning matrix takes vertices from model to joint and back to model space, I'd think the vertices should already be premultiplied by the mesh transform. But if I do so, I just get a black screen. As far as exporting the joints from Blender, my python script exports for each armature bone in bind pose, it's matrix in this way: def DFSJointTraversal(file, skeleton, jointList): for joint in jointList: poseJoint = skeleton.pose.bones[joint.name] jointTransform = poseJoint.matrix.inverted() file.write('Joint ' + joint.name + ' Transform {\n') for col in jointTransform.col: file.write('{:9f} {:9f} {:9f} {:9f}\n'.format(col[0], col[1], col[2], col[3])) DFSJointTraversal(file, skeleton, joint.children) file.write('}\n') And for current / keyframe poses (assuming I'm in the right keyframe): def exportAnimations(filepath): # Only one skeleton per scene objList = [object for object in bpy.context.scene.objects if object.type == 'ARMATURE'] if len(objList) == 0: return elif len(objList) > 1: return #raise exception? dialog box? skeleton = objList[0] jointNames = [bone.name for bone in skeleton.data.bones] for action in bpy.data.actions: # One animation clip per action in Blender, named as the action animationClipFilePath = filepath[0 : filepath.rindex('/') + 1] + action.name + ".aClip" file = open(animationClipFilePath, 'w') file.write('target skeleton: ' + skeleton.name + '\n') file.write('joints count: {:d}'.format(len(jointNames)) + '\n') skeleton.animation_data.action = action keyframeNum = max([len(fcurve.keyframe_points) for fcurve in action.fcurves]) keyframes = [] for fcurve in action.fcurves: for keyframe in fcurve.keyframe_points: keyframes.append(keyframe.co[0]) keyframes = set(keyframes) keyframes = [kf for kf in keyframes] keyframes.sort() file.write('keyframes count: {:d}'.format(len(keyframes)) + '\n') for kfIndex in keyframes: bpy.context.scene.frame_set(kfIndex) file.write('keyframe: {:d}\n'.format(int(kfIndex))) for i in range(0, len(skeleton.data.bones)): file.write('joint: {:d}\n'.format(i)) joint = skeleton.pose.bones[i] jointCurrentPoseTransform = joint.matrix translationV = jointCurrentPoseTransform.to_translation() rotationQ = jointCurrentPoseTransform.to_3x3().to_quaternion() scaleV = jointCurrentPoseTransform.to_scale() file.write('T {:9f} {:9f} {:9f}\n'.format(translationV[0], translationV[1], translationV[2])) file.write('Q {:9f} {:9f} {:9f} {:9f}\n'.format(rotationQ[1], rotationQ[2], rotationQ[3], rotationQ[0])) file.write('S {:9f} {:9f} {:9f}\n'.format(scaleV[0], scaleV[1], scaleV[2])) file.write('\n') file.close() Which I believe follow the theory explained at the beginning of my question. But then I checked out Blender's directX .x exporter for reference.. and what threw me off was that in the .x script they are exporting bind poses like so (transcribed using the same variable names I used so you can compare): if joint.parent: jointTransform = poseJoint.parent.matrix.inverted() else: jointTransform = Matrix() jointTransform *= poseJoint.matrix and exporting current keyframe poses like this: if joint.parent: jointCurrentPoseTransform = joint.parent.matrix.inverted() else: jointCurrentPoseTransform = Matrix() jointCurrentPoseTransform *= joint.matrix why are they using the parent's transform instead of the joint in question's? isn't the join transform assumed to exist in the context of a parent transform since after all it transforms from this joint's space to its parent's? Why are they concatenating in the same order for both bind poses and keyframe poses? If these two are then supposed to be concatenated with each other to cancel out the change of basis? Anyway, any ideas are appreciated.

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  • Can static methods be called using object/instance in .NET

    Ans is Yes and No   Yes in C++, Java and VB.NET No in C#   This is only compiler restriction in c#. You might see in some websites that we can break this restriction using reflection and delegates, but we can’t, according to my little research J I shall try to explain you…   Following is code sample to break this rule using reflection, it seems that it is possible to call a static method using an object, p1 using System; namespace T {     class Program     {         static void Main()         {             var p1 = new Person() { Name = "Smith" };             typeof(Person).GetMethod("TestStatMethod").Invoke(p1, new object[] { });                     }         class Person         {             public string Name { get; set; }             public static void TestStatMethod()             {                 Console.WriteLine("Hello");             }         }     } } but I do not think so this method is being called using p1 rather Type Name “Person”. I shall try to prove this… look at another example…  Test2 has been inherited from Test1. Let’s see various scenarios… Scenario1 using System; namespace T {     class Program     {         static void Main()         {             Test1 t = new Test1();            typeof(Test2).GetMethod("Method1").Invoke(t,                                  new object[] { });         }     }     class Test1     {         public static void Method1()         {             Console.WriteLine("At test1::Method1");         }     }       class Test2 : Test1     {         public static void Method1()         {             Console.WriteLine("At test1::Method2");         }     } } Output:   At test1::Method2 Scenario2         static void Main()         {             Test2 t = new Test2();            typeof(Test2).GetMethod("Method1").Invoke(t,                                          new object[] { });         }   Output:   At test1::Method2   Scenario3         static void Main()         {             Test1 t = new Test2();            typeof(Test2).GetMethod("Method1").Invoke(t,                             new object[] { });         }   Output: At test1::Method2 In all above scenarios output is same, that means, Reflection also not considering the object what you pass to Invoke method in case of static methods. It is always considering the type which you specify in typeof(). So, what is the use passing instance to “Invoke”. Let see below sample using System; namespace T {     class Program     {         static void Main()         {            typeof(Test2).GetMethod("Method1").                Invoke(null, new object[] { });         }     }       class Test1     {         public static void Method1()         {             Console.WriteLine("At test1::Method1");         }     }     class Test2 : Test1     {         public static void Method1()         {             Console.WriteLine("At test1::Method2");         }     } }   Output is   At test1::Method2   I was able to call Invoke “Method1” of Test2 without any object.  Yes, there no wonder here as Method1 is static. So we may conclude that static methods cannot be called using instances (only in c#) Why Microsoft has restricted it in C#? Ans: Really there Is no use calling static methods using objects because static methods are stateless. but still Java and C++ latest compilers allow calling static methods using instances. Java sample class Test {      public static void main(String str[])      {            Person p = new Person();            System.out.println(p.GetCount());      } }   class Person {   public static int GetCount()   {      return 100;   } }   Output          100 span.fullpost {display:none;}

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  • Following the Thread in OSB

    - by Antony Reynolds
    Threading in OSB The Scenario I recently led an OSB POC where we needed to get high throughput from an OSB pipeline that had the following logic: 1. Receive Request 2. Send Request to External System 3. If Response has a particular value   3.1 Modify Request   3.2 Resend Request to External System 4. Send Response back to Requestor All looks very straightforward and no nasty wrinkles along the way.  The flow was implemented in OSB as follows (see diagram for more details): Proxy Service to Receive Request and Send Response Request Pipeline   Copies Original Request for use in step 3 Route Node   Sends Request to External System exposed as a Business Service Response Pipeline   Checks Response to Check If Request Needs to Be Resubmitted Modify Request Callout to External System (same Business Service as Route Node) The Proxy and the Business Service were each assigned their own Work Manager, effectively giving each of them their own thread pool. The Surprise Imagine our surprise when, on stressing the system we saw it lock up, with large numbers of blocked threads.  The reason for the lock up is due to some subtleties in the OSB thread model which is the topic of this post.   Basic Thread Model OSB goes to great lengths to avoid holding on to threads.  Lets start by looking at how how OSB deals with a simple request/response routing to a business service in a route node. Most Business Services are implemented by OSB in two parts.  The first part uses the request thread to send the request to the target.  In the diagram this is represented by the thread T1.  After sending the request to the target (the Business Service in our diagram) the request thread is released back to whatever pool it came from.  A multiplexor (muxer) is used to wait for the response.  When the response is received the muxer hands off the response to a new thread that is used to execute the response pipeline, this is represented in the diagram by T2. OSB allows you to assign different Work Managers and hence different thread pools to each Proxy Service and Business Service.  In out example we have the “Proxy Service Work Manager” assigned to the Proxy Service and the “Business Service Work Manager” assigned to the Business Service.  Note that the Business Service Work Manager is only used to assign the thread to process the response, it is never used to process the request. This architecture means that while waiting for a response from a business service there are no threads in use, which makes for better scalability in terms of thread usage. First Wrinkle Note that if the Proxy and the Business Service both use the same Work Manager then there is potential for starvation.  For example: Request Pipeline makes a blocking callout, say to perform a database read. Business Service response tries to allocate a thread from thread pool but all threads are blocked in the database read. New requests arrive and contend with responses arriving for the available threads. Similar problems can occur if the response pipeline blocks for some reason, maybe a database update for example. Solution The solution to this is to make sure that the Proxy and Business Service use different Work Managers so that they do not contend with each other for threads. Do Nothing Route Thread Model So what happens if there is no route node?  In this case OSB just echoes the Request message as a Response message, but what happens to the threads?  OSB still uses a separate thread for the response, but in this case the Work Manager used is the Default Work Manager. So this is really a special case of the Basic Thread Model discussed above, except that the response pipeline will always execute on the Default Work Manager.   Proxy Chaining Thread Model So what happens when the route node is actually calling a Proxy Service rather than a Business Service, does the second Proxy Service use its own Thread or does it re-use the thread of the original Request Pipeline? Well as you can see from the diagram when a route node calls another proxy service then the original Work Manager is used for both request pipelines.  Similarly the response pipeline uses the Work Manager associated with the ultimate Business Service invoked via a Route Node.  This actually fits in with the earlier description I gave about Business Services and by extension Route Nodes they “… uses the request thread to send the request to the target”. Call Out Threading Model So what happens when you make a Service Callout to a Business Service from within a pipeline.  The documentation says that “The pipeline processor will block the thread until the response arrives asynchronously” when using a Service Callout.  What this means is that the target Business Service is called using the pipeline thread but the response is also handled by the pipeline thread.  This implies that the pipeline thread blocks waiting for a response.  It is the handling of this response that behaves in an unexpected way. When a Business Service is called via a Service Callout, the calling thread is suspended after sending the request, but unlike the Route Node case the thread is not released, it waits for the response.  The muxer uses the Business Service Work Manager to allocate a thread to process the response, but in this case processing the response means getting the response and notifying the blocked pipeline thread that the response is available.  The original pipeline thread can then continue to process the response. Second Wrinkle This leads to an unfortunate wrinkle.  If the Business Service is using the same Work Manager as the Pipeline then it is possible for starvation or a deadlock to occur.  The scenario is as follows: Pipeline makes a Callout and the thread is suspended but still allocated Multiple Pipeline instances using the same Work Manager are in this state (common for a system under load) Response comes back but all Work Manager threads are allocated to blocked pipelines. Response cannot be processed and so pipeline threads never unblock – deadlock! Solution The solution to this is to make sure that any Business Services used by a Callout in a pipeline use a different Work Manager to the pipeline itself. The Solution to My Problem Looking back at my original workflow we see that the same Business Service is called twice, once in a Routing Node and once in a Response Pipeline Callout.  This was what was causing my problem because the response pipeline was using the Business Service Work Manager, but the Service Callout wanted to use the same Work Manager to handle the responses and so eventually my Response Pipeline hogged all the available threads so no responses could be processed. The solution was to create a second Business Service pointing to the same location as the original Business Service, the only difference was to assign a different Work Manager to this Business Service.  This ensured that when the Service Callout completed there were always threads available to process the response because the response processing from the Service Callout had its own dedicated Work Manager. Summary Request Pipeline Executes on Proxy Work Manager (WM) Thread so limited by setting of that WM.  If no WM specified then uses WLS default WM. Route Node Request sent using Proxy WM Thread Proxy WM Thread is released before getting response Muxer is used to handle response Muxer hands off response to Business Service (BS) WM Response Pipeline Executes on Routed Business Service WM Thread so limited by setting of that WM.  If no WM specified then uses WLS default WM. No Route Node (Echo functionality) Proxy WM thread released New thread from the default WM used for response pipeline Service Callout Request sent using proxy pipeline thread Proxy thread is suspended (not released) until the response comes back Notification of response handled by BS WM thread so limited by setting of that WM.  If no WM specified then uses WLS default WM. Note this is a very short lived use of the thread After notification by callout BS WM thread that thread is released and execution continues on the original pipeline thread. Route/Callout to Proxy Service Request Pipeline of callee executes on requestor thread Response Pipeline of caller executes on response thread of requested proxy Throttling Request message may be queued if limit reached. Requesting thread is released (route node) or suspended (callout) So what this means is that you may get deadlocks caused by thread starvation if you use the same thread pool for the business service in a route node and the business service in a callout from the response pipeline because the callout will need a notification thread from the same thread pool as the response pipeline.  This was the problem we were having. You get a similar problem if you use the same work manager for the proxy request pipeline and a business service callout from that request pipeline. It also means you may want to have different work managers for the proxy and business service in the route node. Basically you need to think carefully about how threading impacts your proxy services. References Thanks to Jay Kasi, Gerald Nunn and Deb Ayers for helping to explain this to me.  Any errors are my own and not theirs.  Also thanks to my colleagues Milind Pandit and Prasad Bopardikar who travelled this road with me. OSB Thread Model Great Blog Post on Thread Usage in OSB

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  • In Which We Demystify A Few Docupresentment Settings And Learn the Ethos of the Author

    - by Andy Little
    It's no secret that Docupresentment (part of the Oracle Documaker suite) is powerful tool for integrating on-demand and interactive applications for publishing with the Oracle Documaker framework.  It's also no secret there are are many details with respect to the configuration of Docupresentment that can elude even the most erudite of of techies.  To be sure, Docupresentment will work for you right out of the box, and in most cases will suit your needs without toying with a configuration file.  But, where's the adventure in that?   With this inaugural post to That's The Way, I'm going to introduce myself, and what my aim is with this blog.  If you didn't figure it out already by checking out my profile, my name is Andy and I've been with Oracle (nee Skywire Software nee Docucorp nee Formmaker) since the formative years of 1998.  Strangely, it doesn't seem that long ago, but it's certainly a lifetime in the age of technology.  I recall running a BBS from my parent's basement on a 1200 baud modem, and the trepidation and sweaty-palmed excitement of upgrading to the power and speed of 2400 baud!  Fine, I'll admit that perhaps I'm inflating the experience a bit, but I was kid!  This is the stuff of War Games and King's Quest I and the demise of TI-99 4/A.  Exciting times.  So fast-forward a bit and I'm 12 years into a career in the world of document automation and publishing working for the best (IMHO) software company on the planet.  With That's The Way I hope to shed a little light and peek under the covers of some of the more interesting aspects of implementations involving the tech space within the Oracle Insurance Global Business Unit (IGBU), which includes Oracle Documaker, Rating & Underwriting, and Policy Administration to name a few.  I may delve off course a bit, and you'll likely get a dose of humor (at least in my mind) but I hope you'll glean at least a tidbit of usefulness with each post.  Feel free to comment as I'm a fairly conversant guy and happy to talk -- it's stopping the talking that's the hard part... So, back to our regularly-scheduled post, already in progress.  By this time you've visited Oracle's E-Delivery site and acquired your properly-licensed version of Oracle Documaker.  Wait -- you didn't find it?  Understandable -- navigating the voluminous download library within Oracle can be a daunting task.  It's pretty simple once you’ve done it a few times.  Login to the e-delivery site, and accept the license terms and restrictions.  Then, you’ll be able to select the Oracle Insurance Applications product pack and your appropriate platform. Click Go and you’ll see a list of applicable products, and you’ll click on Oracle Documaker Media Pack (as I went to press with this article the version is 11.4): Finally, click the Download button next to Docupresentment (again, version at press time is 2.2 p5). This should give you a ZIP file that contains the installation packages for the Docupresentment Server and Client, cryptically named IDSServer22P05W32.exe and IDSClient22P05W32.exe. At this time, I’d like to take a little detour and explain that the world of Oracle, like most technical companies, is rife with acronyms.  One of the reasons Skywire Software was a appealing to Oracle was our use of many acronyms, including the occasional use of multiple acronyms with the same meaning.  I apologize in advance and will try to point these out along the way.  Here’s your first sticky note to go along with that: IDS = Internet Document Server = Docupresentment Once you’ve completed the installation, you’ll have a shiny new Docupresentment server and client, and if you installed the default location it will be living in c:\docserv. Unix users, I’m one of you!  You’ll find it by default in  ~/docupresentment/docserv.  Forging onward with the meat of this post is learning about some special configuration options.  By now you’ve read the documentation included with the download (specifically ids_book.pdf) which goes into some detail of the rubric of the configuration file and in fact there’s even a handy utility that provides an interface to the configuration file (see Running IDSConfig in the documentation).  But who wants to deal with a configuration utility when we have the tools and technology to edit the file <gasp> by hand! I shall now proceed with the standard Information Technology Under the Hood Disclaimer: Please remember to back up any files before you make changes.  I am not responsible for any havoc you may wreak! Go to your installation directory, and locate your docserv.xml file.  Open it in your favorite XML editor.  I happen to be fond of Notepad++ with the XML Tools plugin.  Almost immediately you will behold the splendor of the configuration file.  Just take a moment and let that sink in.  Ok – moving on.  If you reviewed the documentation you know that inside the root <configuration> node there are multiple <section> nodes, each containing a specific group of settings.  Let’s take a look at <section name=”DocumentServer”>: There are a few entries I’d like to discuss.  First, <entry name=”StartCommand”>. This should be pretty self-explanatory; it’s the name of the executable that’s run when you fire up Docupresentment.  Immediately following that is <entry name=”StartArguments”> and as you might imagine these are the arguments passed to the executable.  A few things to point out: The –Dids.configuration=docserv.xml parameter specifies the name of your configuration file. The –Dlogging.configuration=logconf.xml parameter specifies the name of your logging configuration file (this uses log4j so bone up on that before you delve here). The -Djava.endorsed.dirs=lib/endorsed parameter specifies the path where 3rd party Java libraries can be located for use with Docupresentment.  More on that in another post. The <entry name=”Instances”> allows you to specify the number of instances of Docupresentment that will be started.  By default this is two, and generally two instances per CPU is adequate, however you will always need to perform load testing to determine the sweet spot based on your hardware and types of transactions.  You may have many, many more instances than 2. Time for a sidebar on instances.  An instance is nothing more than a separate process of Docupresentment.  The Docupresentment service that you fire up with docserver.bat or docserver.sh actually starts a watchdog process, which is then responsible for starting up the actual Docupresentment processes.  Each of these act independently from one another, so if one crashes, it does not affect any others.  In the case of a crashed process, the watchdog will start up another instance so the number of configured instances are always running.  Bottom line: instance = Docupresentment process. And now, finally, to the settings which gave me pause on an not-too-long-ago implementation!  Docupresentment includes a feature that watches configuration files (such as docserv.xml and logconf.xml) and will automatically restart its instances to load the changes.  You can configure the time that Docupresentment waits to check these files using the setting <entry name=”FileWatchTimeMillis”>.  By default the number is 12000ms, or 12 seconds.  You can save yourself a few CPU cycles by extending this time, or by disabling  the check altogether by setting the value to 0.  This may or may not be appropriate for your environment; if you have 100% uptime requirements then you probably don’t want to bring down an entire set of processes just to accept a new configuration value, so it’s best to leave this somewhere between 12 seconds to a few minutes.  Another point to keep in mind: if you are using Documaker real-time processing under Docupresentment the Master Resource Library (MRL) files and INI options are cached, and if you need to affect a change, you’ll have to “restart” Docupresentment.  Touching the docserv.xml file is an easy way to do this (other methods including using the RSS request, but that’s another post). The next item up: <entry name=”FilePurgeTimeSeconds”>.  You may already know that the Docupresentment system can generate many temporary files based on certain request types that are processed through the system.  What you may not know is how those files are cleaned up.  There are many rules in Docupresentment that cause the creation of temporary files.  When these files are created, Docupresentment writes an entry into a properties file called the file cache.  This file contains the name, creation date, and expiration time of each temporary file created by each instance of Docupresentment.  Periodically Docupresentment will check the file cache to determine if there are files that are past the expiration time, not unlike that block of cheese festering away in the back of my refrigerator.  However, unlike my ‘fridge cleaning tendencies, Docupresentment is quick to remove files that are past their expiration time.  You, my friend, have the power to control how often Docupresentment inspects the file cache.  Simply set the value for <entry name=”FilePurgeTimeSeconds”> to the number of seconds appropriate for your requirements and you’re set.  Note that file purging happens on a separate thread from normal request processing, so this shouldn’t interfere with response times unless the CPU happens to be really taxed at the point of cache processing.  Finally, after all of this, we get to the final setting I’m going to address in this post: <entry name=”FilePurgeList”>.  The default is “filecache.properties”.  This establishes the root name for the Docupresentment file cache that I mentioned previously.  Docupresentment creates a separate cache file for each instance based on this setting.  If you have two instances, you’ll see two files created: filecache.properties.1 and filecache.properties.2.  Feel free to open these up and check them out. I hope you’ve enjoyed this first foray into the configuration file of Docupresentment.  If you did enjoy it, feel free to drop a comment, I welcome feedback.  If you have ideas for other posts you’d like to see, please do let me know.  You can reach me at [email protected]. ‘Til next time! ###

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  • What's up with LDoms: Part 1 - Introduction & Basic Concepts

    - by Stefan Hinker
    LDoms - the correct name is Oracle VM Server for SPARC - have been around for quite a while now.  But to my surprise, I get more and more requests to explain how they work or to give advise on how to make good use of them.  This made me think that writing up a few articles discussing the different features would be a good idea.  Now - I don't intend to rewrite the LDoms Admin Guide or to copy and reformat the (hopefully) well known "Beginners Guide to LDoms" by Tony Shoumack from 2007.  Those documents are very recommendable - especially the Beginners Guide, although based on LDoms 1.0, is still a good place to begin with.  However, LDoms have come a long way since then, and I hope to contribute to their adoption by discussing how they work and what features there are today.  In this and the following posts, I will use the term "LDoms" as a common abbreviation for Oracle VM Server for SPARC, just because it's a lot shorter and easier to type (and presumably, read). So, just to get everyone on the same baseline, lets briefly discuss the basic concepts of virtualization with LDoms.  LDoms make use of a hypervisor as a layer of abstraction between real, physical hardware and virtual hardware.  This virtual hardware is then used to create a number of guest systems which each behave very similar to a system running on bare metal:  Each has its own OBP, each will install its own copy of the Solaris OS and each will see a certain amount of CPU, memory, disk and network resources available to it.  Unlike some other type 1 hypervisors running on x86 hardware, the SPARC hypervisor is embedded in the system firmware and makes use both of supporting functions in the sun4v SPARC instruction set as well as the overall CPU architecture to fulfill its function. The CMT architecture of the supporting CPUs (T1 through T4) provide a large number of cores and threads to the OS.  For example, the current T4 CPU has eight cores, each running 8 threads, for a total of 64 threads per socket.  To the OS, this looks like 64 CPUs.  The SPARC hypervisor, when creating guest systems, simply assigns a certain number of these threads exclusively to one guest, thus avoiding the overhead of having to schedule OS threads to CPUs, as do typical x86 hypervisors.  The hypervisor only assigns CPUs and then steps aside.  It is not involved in the actual work being dispatched from the OS to the CPU, all it does is maintain isolation between different guests. Likewise, memory is assigned exclusively to individual guests.  Here,  the hypervisor provides generic mappings between the physical hardware addresses and the guest's views on memory.  Again, the hypervisor is not involved in the actual memory access, it only maintains isolation between guests. During the inital setup of a system with LDoms, you start with one special domain, called the Control Domain.  Initially, this domain owns all the hardware available in the system, including all CPUs, all RAM and all IO resources.  If you'd be running the system un-virtualized, this would be what you'd be working with.  To allow for guests, you first resize this initial domain (also called a primary domain in LDoms speak), assigning it a small amount of CPU and memory.  This frees up most of the available CPU and memory resources for guest domains.  IO is a little more complex, but very straightforward.  When LDoms 1.0 first came out, the only way to provide IO to guest systems was to create virtual disk and network services and attach guests to these services.  In the meantime, several different ways to connect guest domains to IO have been developed, the most recent one being SR-IOV support for network devices released in version 2.2 of Oracle VM Server for SPARC. I will cover these more advanced features in detail later.  For now, lets have a short look at the initial way IO was virtualized in LDoms: For virtualized IO, you create two services, one "Virtual Disk Service" or vds, and one "Virtual Switch" or vswitch.  You can, of course, also create more of these, but that's more advanced than I want to cover in this introduction.  These IO services now connect real, physical IO resources like a disk LUN or a networt port to the virtual devices that are assigned to guest domains.  For disk IO, the normal case would be to connect a physical LUN (or some other storage option that I'll discuss later) to one specific guest.  That guest would be assigned a virtual disk, which would appear to be just like a real LUN to the guest, while the IO is actually routed through the virtual disk service down to the physical device.  For network, the vswitch acts very much like a real, physical ethernet switch - you connect one physical port to it for outside connectivity and define one or more connections per guest, just like you would plug cables between a real switch and a real system. For completeness, there is another service that provides console access to guest domains which mimics the behavior of serial terminal servers. The connections between the virtual devices on the guest's side and the virtual IO services in the primary domain are created by the hypervisor.  It uses so called "Logical Domain Channels" or LDCs to create point-to-point connections between all of these devices and services.  These LDCs work very similar to high speed serial connections and are configured automatically whenever the Control Domain adds or removes virtual IO. To see all this in action, now lets look at a first example.  I will start with a newly installed machine and configure the control domain so that it's ready to create guest systems. In a first step, after we've installed the software, let's start the virtual console service and downsize the primary domain.  root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-c-- UART 512 261632M 0.3% 2d 13h 58m root@sun # ldm add-vconscon port-range=5000-5100 \ primary-console primary root@sun # svcadm enable vntsd root@sun # svcs vntsd STATE STIME FMRI online 9:53:21 svc:/ldoms/vntsd:default root@sun # ldm set-vcpu 16 primary root@sun # ldm set-mau 1 primary root@sun # ldm start-reconf primary root@sun # ldm set-memory 7680m primary root@sun # ldm add-config initial root@sun # shutdown -y -g0 -i6 So what have I done: I've defined a range of ports (5000-5100) for the virtual network terminal service and then started that service.  The vnts will later provide console connections to guest systems, very much like serial NTS's do in the physical world. Next, I assigned 16 vCPUs (on this platform, a T3-4, that's two cores) to the primary domain, freeing the rest up for future guest systems.  I also assigned one MAU to this domain.  A MAU is a crypto unit in the T3 CPU.  These need to be explicitly assigned to domains, just like CPU or memory.  (This is no longer the case with T4 systems, where crypto is always available everywhere.) Before I reassigned the memory, I started what's called a "delayed reconfiguration" session.  That avoids actually doing the change right away, which would take a considerable amount of time in this case.  Instead, I'll need to reboot once I'm all done.  I've assigned 7680MB of RAM to the primary.  That's 8GB less the 512MB which the hypervisor uses for it's own private purposes.  You can, depending on your needs, work with less.  I'll spend a dedicated article on sizing, discussing the pros and cons in detail. Finally, just before the reboot, I saved my work on the ILOM, to make this configuration available after a powercycle of the box.  (It'll always be available after a simple reboot, but the ILOM needs to know the configuration of the hypervisor after a power-cycle, before the primary domain is booted.) Now, lets create a first disk service and a first virtual switch which is connected to the physical network device igb2. We will later use these to connect virtual disks and virtual network ports of our guest systems to real world storage and network. root@sun # ldm add-vds primary-vds root@sun # ldm add-vswitch net-dev=igb2 switch-primary primary You are free to choose whatever names you like for the virtual disk service and the virtual switch.  I strongly recommend that you choose names that make sense to you and describe the function of each service in the context of your implementation.  For the vswitch, for example, you could choose names like "admin-vswitch" or "production-network" etc. This already concludes the configuration of the control domain.  We've freed up considerable amounts of CPU and RAM for guest systems and created the necessary infrastructure - console, vts and vswitch - so that guests systems can actually interact with the outside world.  The system is now ready to create guests, which I'll describe in the next section. For further reading, here are some recommendable links: The LDoms 2.2 Admin Guide The "Beginners Guide to LDoms" The LDoms Information Center on MOS LDoms on OTN

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  • Performance Optimization &ndash; It Is Faster When You Can Measure It

    - by Alois Kraus
    Performance optimization in bigger systems is hard because the measured numbers can vary greatly depending on the measurement method of your choice. To measure execution timing of specific methods in your application you usually use Time Measurement Method Potential Pitfalls Stopwatch Most accurate method on recent processors. Internally it uses the RDTSC instruction. Since the counter is processor specific you can get greatly different values when your thread is scheduled to another core or the core goes into a power saving mode. But things do change luckily: Intel's Designer's vol3b, section 16.11.1 "16.11.1 Invariant TSC The time stamp counter in newer processors may support an enhancement, referred to as invariant TSC. Processor's support for invariant TSC is indicated by CPUID.80000007H:EDX[8]. The invariant TSC will run at a constant rate in all ACPI P-, C-. and T-states. This is the architectural behavior moving forward. On processors with invariant TSC support, the OS may use the TSC for wall clock timer services (instead of ACPI or HPET timers). TSC reads are much more efficient and do not incur the overhead associated with a ring transition or access to a platform resource." DateTime.Now Good but it has only a resolution of 16ms which can be not enough if you want more accuracy.   Reporting Method Potential Pitfalls Console.WriteLine Ok if not called too often. Debug.Print Are you really measuring performance with Debug Builds? Shame on you. Trace.WriteLine Better but you need to plug in some good output listener like a trace file. But be aware that the first time you call this method it will read your app.config and deserialize your system.diagnostics section which does also take time.   In general it is a good idea to use some tracing library which does measure the timing for you and you only need to decorate some methods with tracing so you can later verify if something has changed for the better or worse. In my previous article I did compare measuring performance with quantum mechanics. This analogy does work surprising well. When you measure a quantum system there is a lower limit how accurately you can measure something. The Heisenberg uncertainty relation does tell us that you cannot measure of a quantum system the impulse and location of a particle at the same time with infinite accuracy. For programmers the two variables are execution time and memory allocations. If you try to measure the timings of all methods in your application you will need to store them somewhere. The fastest storage space besides the CPU cache is the memory. But if your timing values do consume all available memory there is no memory left for the actual application to run. On the other hand if you try to record all memory allocations of your application you will also need to store the data somewhere. This will cost you memory and execution time. These constraints are always there and regardless how good the marketing of tool vendors for performance and memory profilers are: Any measurement will disturb the system in a non predictable way. Commercial tool vendors will tell you they do calculate this overhead and subtract it from the measured values to give you the most accurate values but in reality it is not entirely true. After falling into the trap to trust the profiler timings several times I have got into the habit to Measure with a profiler to get an idea where potential bottlenecks are. Measure again with tracing only the specific methods to check if this method is really worth optimizing. Optimize it Measure again. Be surprised that your optimization has made things worse. Think harder Implement something that really works. Measure again Finished! - Or look for the next bottleneck. Recently I have looked into issues with serialization performance. For serialization DataContractSerializer was used and I was not sure if XML is really the most optimal wire format. After looking around I have found protobuf-net which uses Googles Protocol Buffer format which is a compact binary serialization format. What is good for Google should be good for us. A small sample app to check out performance was a matter of minutes: using ProtoBuf; using System; using System.Diagnostics; using System.IO; using System.Reflection; using System.Runtime.Serialization; [DataContract, Serializable] class Data { [DataMember(Order=1)] public int IntValue { get; set; } [DataMember(Order = 2)] public string StringValue { get; set; } [DataMember(Order = 3)] public bool IsActivated { get; set; } [DataMember(Order = 4)] public BindingFlags Flags { get; set; } } class Program { static MemoryStream _Stream = new MemoryStream(); static MemoryStream Stream { get { _Stream.Position = 0; _Stream.SetLength(0); return _Stream; } } static void Main(string[] args) { DataContractSerializer ser = new DataContractSerializer(typeof(Data)); Data data = new Data { IntValue = 100, IsActivated = true, StringValue = "Hi this is a small string value to check if serialization does work as expected" }; var sw = Stopwatch.StartNew(); int Runs = 1000 * 1000; for (int i = 0; i < Runs; i++) { //ser.WriteObject(Stream, data); Serializer.Serialize<Data>(Stream, data); } sw.Stop(); Console.WriteLine("Did take {0:N0}ms for {1:N0} objects", sw.Elapsed.TotalMilliseconds, Runs); Console.ReadLine(); } } The results are indeed promising: Serializer Time in ms N objects protobuf-net   807 1000000 DataContract 4402 1000000 Nearly a factor 5 faster and a much more compact wire format. Lets use it! After switching over to protbuf-net the transfered wire data has dropped by a factor two (good) and the performance has worsened by nearly a factor two. How is that possible? We have measured it? Protobuf-net is much faster! As it turns out protobuf-net is faster but it has a cost: For the first time a type is de/serialized it does use some very smart code-gen which does not come for free. Lets try to measure this one by setting of our performance test app the Runs value not to one million but to 1. Serializer Time in ms N objects protobuf-net 85 1 DataContract 24 1 The code-gen overhead is significant and can take up to 200ms for more complex types. The break even point where the code-gen cost is amortized by its faster serialization performance is (assuming small objects) somewhere between 20.000-40.000 serialized objects. As it turned out my specific scenario involved about 100 types and 1000 serializations in total. That explains why the good old DataContractSerializer is not so easy to take out of business. The final approach I ended up was to reduce the number of types and to serialize primitive types via BinaryWriter directly which turned out to be a pretty good alternative. It sounded good until I measured again and found that my optimizations so far do not help much. After looking more deeper at the profiling data I did found that one of the 1000 calls did take 50% of the time. So how do I find out which call it was? Normal profilers do fail short at this discipline. A (totally undeserved) relatively unknown profiler is SpeedTrace which does unlike normal profilers create traces of your applications by instrumenting your IL code at runtime. This way you can look at the full call stack of the one slow serializer call to find out if this stack was something special. Unfortunately the call stack showed nothing special. But luckily I have my own tracing as well and I could see that the slow serializer call did happen during the serialization of a bool value. When you encounter after much analysis something unreasonable you cannot explain it then the chances are good that your thread was suspended by the garbage collector. If there is a problem with excessive GCs remains to be investigated but so far the serialization performance seems to be mostly ok.  When you do profile a complex system with many interconnected processes you can never be sure that the timings you just did measure are accurate at all. Some process might be hitting the disc slowing things down for all other processes for some seconds as well. There is a big difference between warm and cold startup. If you restart all processes you can basically forget the first run because of the OS disc cache, JIT and GCs make the measured timings very flexible. When you are in need of a random number generator you should measure cold startup times of a sufficiently complex system. After the first run you can try again getting different and much lower numbers. Now try again at least two times to get some feeling how stable the numbers are. Oh and try to do the same thing the next day. It might be that the bottleneck you found yesterday is gone today. Thanks to GC and other random stuff it can become pretty hard to find stuff worth optimizing if no big bottlenecks except bloatloads of code are left anymore. When I have found a spot worth optimizing I do make the code changes and do measure again to check if something has changed. If it has got slower and I am certain that my change should have made it faster I can blame the GC again. The thing is that if you optimize stuff and you allocate less objects the GC times will shift to some other location. If you are unlucky it will make your faster working code slower because you see now GCs at times where none were before. This is where the stuff does get really tricky. A safe escape hatch is to create a repro of the slow code in an isolated application so you can change things fast in a reliable manner. Then the normal profilers do also start working again. As Vance Morrison does point out it is much more complex to profile a system against the wall clock compared to optimize for CPU time. The reason is that for wall clock time analysis you need to understand how your system does work and which threads (if you have not one but perhaps 20) are causing a visible delay to the end user and which threads can wait a long time without affecting the user experience at all. Next time: Commercial profiler shootout.

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  • EM12c Release 4: Cloud Control to Major Tom...

    - by abulloch
    With the latest release of Enterprise Manager 12c, Release 4 (12.1.0.4) the EM development team has added new functionality to assist the EM Administrator to monitor the health of the EM infrastructure.   Taking feedback delivered from customers directly and through customer advisory boards some nice enhancements have been made to the “Manage Cloud Control” sections of the UI, commonly known in the EM community as “the MTM pages” (MTM stands for Monitor the Monitor).  This part of the EM Cloud Control UI is viewed by many as the mission control for EM Administrators. In this post we’ll highlight some of the new information that’s on display in these redesigned pages and explain how the information they present can help EM administrators identify potential bottlenecks or issues with the EM infrastructure. The first page we’ll take a look at is the newly designed Repository information page.  You can get to this from the main Setup menu, through Manage Cloud Control, then Repository.  Once this page loads you’ll see the new layout that includes 3 tabs containing more drill-down information. The Repository Tab The first tab, Repository, gives you a series of 6 panels or regions on screen that display key information that the EM Administrator needs to review from time to time to ensure that their infrastructure is in good health. Rather than go through every panel let’s call out a few and let you explore the others later yourself on your own EM site.  Firstly, we have the Repository Details panel. At a glance the EM Administrator can see the current version of the EM repository database and more critically, three important elements of information relating to availability and reliability :- Is the database in Archive Log mode ? Is the database using Flashback ? When was the last database backup taken ? In this test environment above the answers are not too worrying, however, Production environments should have at least Archivelog mode enabled, Flashback is a nice feature to enable prior to upgrades (for fast rollback) and all Production sites should have a backup.  In this case the backup information in the Control file indicates there’s been no recorded backups taken. The next region of interest to note on this page shows key information around the Repository configuration, specifically, the initialisation parameters (from the spfile). If you’re storing your EM Repository in a Cluster Database you can view the parameters on each individual instance using the Instance Name drop-down selector in the top right of the region. Additionally, you’ll note there is now a check performed on the active configuration to ensure that you’re using, at the very least, Oracle minimum recommended values.  Should the values in your EM Repository not meet these requirements it will be flagged in this table with a red X for non-compliance.  You can of-course change these values within EM by selecting the Database target and modifying the parameters in the spfile (and optionally, the run-time values if the parameter allows dynamic changes). The last region to call out on this page before moving on is the new look Repository Scheduler Job Status region. This region is an update of a similar region seen on previous releases of the MTM pages in Cloud Control but there’s some important new functionality that’s been added that customers have requested. First-up - Restarting Repository Jobs.  As you can see from the graphic, you can now optionally select a job (by selecting the row in the UI table element) and click on the Restart Job button to take care of any jobs which have stopped or stalled for any reason.  Previously this needed to be done at the command line using EMDIAG or through a PL/SQL package invocation.  You can now take care of this directly from within the UI. Next, you’ll see that a feature has been added to allow the EM administrator to customise the run-time for some of the background jobs that run in the Repository.  We heard from some customers that ensuring these jobs don’t clash with Production backups, etc is a key requirement.  This new functionality allows you to select the pencil icon to edit the schedule time for these more resource intensive background jobs and modify the schedule to avoid clashes like this. Moving onto the next tab, let’s select the Metrics tab. The Metrics Tab There’s some big changes here, this page contains new information regions that help the Administrator understand the direct impact the in-bound metric flows are having on the EM Repository.  Many customers have provided feedback that they are in the dark about the impact of adding new targets or large numbers of new hosts or new target types into EM and the impact this has on the Repository.  This page helps the EM Administrator get to grips with this.  Let’s take a quick look at two regions on this page. First-up there’s a bubble chart showing a comprehensive view of the top resource consumers of metric data, over the last 30 days, charted as the number of rows loaded against the number of collections for the metric.  The size of the bubble indicates a relative volume.  You can see from this example above that a quick glance shows that Host metrics are the largest inbound flow into the repository when measured by number of rows.  Closely following behind this though are a large number of collections for Oracle Weblogic Server and Application Deployment.  Taken together the Host Collections is around 0.7Mb of data.  The total information collection for Weblogic Server and Application Deployments is 0.38Mb and 0.37Mb respectively. If you want to get this information breakdown on the volume of data collected simply hover over the bubble in the chart and you’ll get a floating tooltip showing the information. Clicking on any bubble in the chart takes you one level deeper into a drill-down of the Metric collection. Doing this reveals the individual metric elements for these target types and again shows a representation of the relative cost - in terms of Number of Rows, Number of Collections and Storage cost of data for each Metric type. Looking at another panel on this page we can see a different view on this data. This view shows a view of the Top N metrics (the drop down allows you to select 10, 15 or 20) and sort them by volume of data.  In the case above we can see the largest metric collection (by volume) in this case (over the last 30 days) is the information about OS Registered Software on a Host target. Taken together, these two regions provide a powerful tool for the EM Administrator to understand the potential impact of any new targets that have been discovered and promoted into management by EM12c.  It’s a great tool for identifying the cause of a sudden increase in Repository storage consumption or Redo log and Archive log generation. Using the information on this page EM Administrators can take action to mitigate any load impact by deploying monitoring templates to the targets causing most load if appropriate.   The last tab we’ll look at on this page is the Schema tab. The Schema Tab Selecting this tab brings up a window onto the SYSMAN schema with a focus on Space usage in the EM Repository.  Understanding what tablespaces are growing, at what rate, is essential information for the EM Administrator to stay on top of managing space allocations for the EM Repository so that it works as efficiently as possible and performs well for the users.  Not least because ensuring storage is managed well ensures continued availability of EM for monitoring purposes. The first region to highlight here shows the trend of space usage for the tablespaces in the EM Repository over time.  You can see the upward trend here showing that storage in the EM Repository is being consumed on an upward trend over the last few days here. This is normal as this EM being used here is brand new with Agents being added daily to bring targets into monitoring.  If your Enterprise Manager configuration has reached a steady state over a period of time where the number of new inbound targets is relatively small, the metric collection settings are fairly uniform and standardised (using Templates and Template Collections) you’re likely to see a trend of space allocation that plateau’s. The table below the trend chart shows the Top 20 Tables/Indexes sorted descending by order of space consumed.  You can switch the trend view chart and corresponding detail table by choosing a different tablespace in the EM Repository using the drop-down picker on the top right of this region. The last region to highlight on this page is the region showing information about the Purge policies in effect in the EM Repository. This information is useful to illustrate to EM Administrators the default purge policies in effect for the different categories of information available in the EM Repository.  Of course, it’s also been a long requested feature to have the ability to modify these default retention periods.  You can also do this using this screen.  As there are interdependencies between some data elements you can’t modify retention policies on a feature by feature basis.  Instead, retention policies take categories of information and bundles them together in Groups.  Retention policies are modified at the Group Level.  Understanding the impact of this really deserves a blog post all on it’s own as modifying these can have a significant impact on both the EM Repository’s storage footprint and it’s performance.  For now, we’re just highlighting the features visibility on these new pages. As a user of EM12c we hope the new features you see here address some of the feedback that’s been given on these pages over the past few releases.  We’ll look out for any comments or feedback you have on these pages ! 

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  • Refactor This (Ugly Code)!

    - by Alois Kraus
    Ayende has put on his blog some ugly code to refactor. First and foremost it is nearly impossible to reason about other peoples code without knowing the driving forces behind the current code. It is certainly possible to make it much cleaner when potential sources of errors cannot happen in the first place due to good design. I can see what the intention of the code is but I do not know about every brittle detail if I am allowed to reorder things here and there to simplify things. So I decided to make it much simpler by identifying the different responsibilities of the methods and encapsulate it in different classes. The code we need to refactor seems to deal with a handler after a message has been sent to a message queue. The handler does complete the current transaction if there is any and does handle any errors happening there. If during the the completion of the transaction errors occur the transaction is at least disposed. We can enter the handler already in a faulty state where we try to deliver the complete event in any case and signal a failure event and try to resend the message again to the queue if it was not inside a transaction. All is decorated with many try/catch blocks, duplicated code and some state variables to route the program flow. It is hard to understand and difficult to reason about. In other words: This code is a mess and could be written by me if I was under pressure. Here comes to code we want to refactor:         private void HandleMessageCompletion(                                      Message message,                                      TransactionScope tx,                                      OpenedQueue messageQueue,                                      Exception exception,                                      Action<CurrentMessageInformation, Exception> messageCompleted,                                      Action<CurrentMessageInformation> beforeTransactionCommit)         {             var txDisposed = false;             if (exception == null)             {                 try                 {                     if (tx != null)                     {                         if (beforeTransactionCommit != null)                             beforeTransactionCommit(currentMessageInformation);                         tx.Complete();                         tx.Dispose();                         txDisposed = true;                     }                     try                     {                         if (messageCompleted != null)                             messageCompleted(currentMessageInformation, exception);                     }                     catch (Exception e)                     {                         Trace.TraceError("An error occured when raising the MessageCompleted event, the error will NOT affect the message processing"+ e);                     }                     return;                 }                 catch (Exception e)                 {                     Trace.TraceWarning("Failed to complete transaction, moving to error mode"+ e);                     exception = e;                 }             }             try             {                 if (txDisposed == false && tx != null)                 {                     Trace.TraceWarning("Disposing transaction in error mode");                     tx.Dispose();                 }             }             catch (Exception e)             {                 Trace.TraceWarning("Failed to dispose of transaction in error mode."+ e);             }             if (message == null)                 return;                 try             {                 if (messageCompleted != null)                     messageCompleted(currentMessageInformation, exception);             }             catch (Exception e)             {                 Trace.TraceError("An error occured when raising the MessageCompleted event, the error will NOT affect the message processing"+ e);             }               try             {                 var copy = MessageProcessingFailure;                 if (copy != null)                     copy(currentMessageInformation, exception);             }             catch (Exception moduleException)             {                 Trace.TraceError("Module failed to process message failure: " + exception.Message+                                              moduleException);             }               if (messageQueue.IsTransactional == false)// put the item back in the queue             {                 messageQueue.Send(message);             }         }     You can see quite some processing and handling going on there. Yes this looks like real world code one did put together to make things work and he does not trust his callbacks. I guess these are event handlers which are optional and the delegates were extracted from an event to call them back later when necessary.  Lets see what the author of this code did intend:          private void HandleMessageCompletion(             TransactionHandler transactionHandler,             MessageCompletionHandler handler,             CurrentMessageInformation messageInfo,             ErrorCollector errors             )         {               // commit current pending transaction             transactionHandler.CallHandlerAndCommit(messageInfo, errors);               // We have an error for a null message do not send completion event             if (messageInfo.CurrentMessage == null)                 return;               // Send completion event in any case regardless of errors             handler.OnMessageCompleted(messageInfo, errors);               // put message back if queue is not transactional             transactionHandler.ResendMessageOnError(messageInfo.CurrentMessage, errors);         }   I did not bother to write the intention here again since the code should be pretty self explaining by now. I have used comments to explain the still nontrivial procedure step by step revealing the real intention about all this complex program flow. The original complexity of the problem domain does not go away but by applying the techniques of SRP (Single Responsibility Principle) and some functional style but we can abstract the necessary complexity away in useful abstractions which make it much easier to reason about it. Since most of the method seems to deal with errors I thought it was a good idea to encapsulate the error state of our current message in an ErrorCollector object which stores all exceptions in a list along with a description what the error all was about in the exception itself. We can log it later or not depending on the log level or whatever. It is really just a simple list that encapsulates the current error state.          class ErrorCollector          {              List<Exception> _Errors = new List<Exception>();                public void Add(Exception ex, string description)              {                  ex.Data["Description"] = description;                  _Errors.Add(ex);              }                public Exception Last              {                  get                  {                      return _Errors.LastOrDefault();                  }              }                public bool HasError              {                  get                  {                      return _Errors.Count > 0;                  }              }          }   Since the error state is global we have two choices to store a reference in the other helper objects (TransactionHandler and MessageCompletionHandler)or pass it to the method calls when necessary. I did chose the latter one because a second argument does not hurt and makes it easier to reason about the overall state while the helper objects remain stateless and immutable which makes the helper objects much easier to understand and as a bonus thread safe as well. This does not mean that the stored member variables are stateless or thread safe as well but at least our helper classes are it. Most of the complexity is located the transaction handling I consider as a separate responsibility that I delegate to the TransactionHandler which does nothing if there is no transaction or Call the Before Commit Handler Commit Transaction Dispose Transaction if commit did throw In fact it has a second responsibility to resend the message if the transaction did fail. I did see a good fit there since it deals with transaction failures.          class TransactionHandler          {              TransactionScope _Tx;              Action<CurrentMessageInformation> _BeforeCommit;              OpenedQueue _MessageQueue;                public TransactionHandler(TransactionScope tx, Action<CurrentMessageInformation> beforeCommit, OpenedQueue messageQueue)              {                  _Tx = tx;                  _BeforeCommit = beforeCommit;                  _MessageQueue = messageQueue;              }                public void CallHandlerAndCommit(CurrentMessageInformation currentMessageInfo, ErrorCollector errors)              {                  if (_Tx != null && !errors.HasError)                  {                      try                      {                          if (_BeforeCommit != null)                          {                              _BeforeCommit(currentMessageInfo);                          }                            _Tx.Complete();                          _Tx.Dispose();                      }                      catch (Exception ex)                      {                          errors.Add(ex, "Failed to complete transaction, moving to error mode");                          Trace.TraceWarning("Disposing transaction in error mode");                          try                          {                              _Tx.Dispose();                          }                          catch (Exception ex2)                          {                              errors.Add(ex2, "Failed to dispose of transaction in error mode.");                          }                      }                  }              }                public void ResendMessageOnError(Message message, ErrorCollector errors)              {                  if (errors.HasError && !_MessageQueue.IsTransactional)                  {                      _MessageQueue.Send(message);                  }              }          } If we need to change the handling in the future we have a much easier time to reason about our application flow than before. After we did complete our transaction and called our callback we can call the completion handler which is the main purpose of the HandleMessageCompletion method after all. The responsiblity o the MessageCompletionHandler is to call the completion callback and the failure callback when some error has occurred.            class MessageCompletionHandler          {              Action<CurrentMessageInformation, Exception> _MessageCompletedHandler;              Action<CurrentMessageInformation, Exception> _MessageProcessingFailure;                public MessageCompletionHandler(Action<CurrentMessageInformation, Exception> messageCompletedHandler,                                              Action<CurrentMessageInformation, Exception> messageProcessingFailure)              {                  _MessageCompletedHandler = messageCompletedHandler;                  _MessageProcessingFailure = messageProcessingFailure;              }                  public void OnMessageCompleted(CurrentMessageInformation currentMessageInfo, ErrorCollector errors)              {                  try                  {                      if (_MessageCompletedHandler != null)                      {                          _MessageCompletedHandler(currentMessageInfo, errors.Last);                      }                  }                  catch (Exception ex)                  {                      errors.Add(ex, "An error occured when raising the MessageCompleted event, the error will NOT affect the message processing");                  }                    if (errors.HasError)                  {                      SignalFailedMessage(currentMessageInfo, errors);                  }              }                void SignalFailedMessage(CurrentMessageInformation currentMessageInfo, ErrorCollector errors)              {                  try                  {                      if (_MessageProcessingFailure != null)                          _MessageProcessingFailure(currentMessageInfo, errors.Last);                  }                  catch (Exception moduleException)                  {                      errors.Add(moduleException, "Module failed to process message failure");                  }              }            }   If for some reason I did screw up the logic and we need to call the completion handler from our Transaction handler we can simple add to the CallHandlerAndCommit method a third argument to the MessageCompletionHandler and we are fine again. If the logic becomes even more complex and we need to ensure that the completed event is triggered only once we have now one place the completion handler to capture the state. During this refactoring I simple put things together that belong together and came up with useful abstractions. If you look at the original argument list of the HandleMessageCompletion method I have put many things together:   Original Arguments New Arguments Encapsulate Message message CurrentMessageInformation messageInfo         Message message TransactionScope tx Action<CurrentMessageInformation> beforeTransactionCommit OpenedQueue messageQueue TransactionHandler transactionHandler        TransactionScope tx        OpenedQueue messageQueue        Action<CurrentMessageInformation> beforeTransactionCommit Exception exception,             ErrorCollector errors Action<CurrentMessageInformation, Exception> messageCompleted MessageCompletionHandler handler          Action<CurrentMessageInformation, Exception> messageCompleted          Action<CurrentMessageInformation, Exception> messageProcessingFailure The reason is simple: Put the things that have relationships together and you will find nearly automatically useful abstractions. I hope this makes sense to you. If you see a way to make it even more simple you can show Ayende your improved version as well.

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  • VSNewFile: A Visual Studio Addin to More Easily Add New Items to a Project

    - by InfinitiesLoop
    My first Visual Studio Add-in! Creating add-ins is pretty simple, once you get used to the CommandBar model it is using, which is apparently a general Office suite extensibility mechanism. Anyway, let me first explain my motivation for this. It started out as an academic exercise, as I have always wanted to dip my feet in a little VS extensibility. But I thought of a legitimate need for an add-in, at least in my personal experience, so it took on new life. But I figured I can’t be the only one who has felt this way, so I decided to publish the add-in, and host it on GitHub (VSNewFile on GitHub) hoping to spur contributions. Adding Files the Built-in Way Here’s the problem I wanted to solve. You’re working on a project, and it’s time to add a new file to the project. Whatever it is – a class, script, html page, aspx page, or what-have-you, you go through a menu or keyboard shortcut to get to the “Add New Item” dialog. Typically, you do it by right-clicking the location where you want the file (the project or a folder of it): This brings up a dialog the contains, well, every conceivable type of item you might want to add. It’s all the available item templates, which can result in anywhere from a ton to a veritable sea of choices. To be fair, this dialog has been revamped in Visual Studio 2010, which organizes it a little better than Visual Studio 2008, and adds a search box. It also loads noticeably faster.   To me, this dialog is just getting in my way. If I want to add a JavaScript script to my project, I don’t want to have to hunt for the script template item in this dialog. Yes, it is categorized, and yes, it now has a search box. But still, all this UI to swim through when all I need is a new file in the project. I will name it. I will provide the content, I don’t even need a ‘template’. VS kind of realizes this. In the add menu in a class library project, for example, there is a “Add Class…” choice. But all this really does is select that project item from the dialog by default. You still must wait for the dialog, see it, and type in a name for the file. How is that really any different than hitting F2 on an existing item? It isn’t. Adding Files the Hack Way What I often find myself doing, just to avoid going through this dialog, is to copy and paste an existing file, rename it, then “CTRL-A, DEL” the content. In a few short keystrokes I’ve got my new file. Even if the original file wasn’t the right type, it doesn’t matter – I will rename it anyway, including the extension. It works well enough if the place I am adding the file to doesn’t have much in it already. But if there are a lot of files at that level, it sucks, because the new file will have the name “Copy of xyz”, causing it to be moved into the ‘C’ section of the alphabetically sorted items, which might be far, far away from the original file (and so I tend to try and copy a file that starts with ‘C’ *evil grin*). Using ‘Export Template’ To be completely fair I should at least mention this feature. I’m not even sure if this is new in VS 2010 or not (I think so). But it allows you to export a project item or items, including potential project references required by it. Then it becomes a new item in the available ‘installed templates’. No doubt this is useful to help bootstrap new projects. But that still requires you to go through the ‘New Item’ dialog. Adding Files with VSNewFile So hopefully I have sufficiently defined the problem and got a few of you to think, “Yeah, me too!”… What VSNewFile does is let you skip the dialog entirely by adding project items directly to the context menu. But it does a bit more than that, so do read on. For example, to add a new class, you can right-click the location and pick that option. A new .cs file is instantly added to the project, and the new item is selected and put into the ‘rename’ mode immediately. The default items available are shown here. But you can customize them. You can also customize the content of each template. To do so, you create a directory in your documents folder, ‘VSNewFile Templates’. In there, you drop the templates you want to use, but you name them in a particular way. For example, here’s a template that will add a new item named “Add TITLE”. It will add a project item named “SOMEFILE.foo” (or ‘SOMEFILE1.foo’ if that exists, etc). The format of the file name is: <ORDER>_<KEY>_<BASE FILENAME>_<ICON ID>_<TITLE>.<EXTENTION> Where: <ORDER> is a number that lets you determine the order of the items in the menu (relative to each other). <KEY> is a case sensitive identifier different for each template item. More on that later. <BASE FILENAME> is the default name of the file, which doesn’t matter that much, since they will be renaming it anyway. <ICON ID> is a number the dictates the icon used for the menu item. There are a huge number of built-in choices. More on that later. <TITLE> is the string that will appear in the menu. And, the contents of the file are the default content for the item (the ‘template’). The content of the file can contain anything you want, of course. But it also supports two tokens: %NAMESPACE% and %FILENAME%, which will be replaced with the corresponding values. Here is the content of this sample: testing Namespace = %NAMESPACE% Filename = %FILENAME% I kind went back and forth on this. I could have made it so there’d be an XML or JSON file that defines the templates, instead of cramming all this data into the filename itself. I like the simplicity of this better. It makes it easy to customize since you can literally just throw these files around, copy them from someone else, etc, without worrying about merge data into a central description file, in whatever format. Here’s our new item showing up: Practical Use One immediate thing I am using this for is to make it easier to add very commonly used scripts to my web projects. For example, uh, say, jQuery? :) All I need to do is drop jQuery-1.4.2.js and jQuery-1.4.2.min.js into the templates folder, provide the order, title, etc, and then instantly, I can now add jQuery to any project I have without even thinking about “where is jQuery? Can I copy it from that other project?”   Using the KEY There are two reasons for the ‘key’ portion of the item. First, it allows you to turn off the built-in, default templates, which are: FILE = Add File (generic, empty file) VB = Add VB Class CS = Add C# Class (includes some basic usings) HTML = Add HTML page (includes basic structure, doctype, etc) JS = Add Script (includes an immediately-invoking function closure) To turn one off, just include a file with the name “_<KEY>”. For example, to turn off all the items except our custom one, you do this: The other reason for the key is that there are new Visual Studio Commands created for each one. This makes it possible to bind a keyboard shortcut to one of them. So you could, for example, have a keyboard combination that adds a new web page to your website, or a new CS class to your class library, etc. Here is our sample item showing up in the keyboard bindings option. Even though the contents of the template directory may change from one launch of Visual Studio to the next, the bindings will remain attached to any item with a particular key, thanks to it taking care not to lose keyboard bindings even though the commands are completely recreated each time. The Icon Face ID Visual Studio uses a Microsoft Office style add-in mechanism, I gather. There are a predetermined set of built-in icons available. You can use your own icons when developing add-ins, of course, but I’m no designer. I just wanted to find appropriate-ish icons for the built-in templates, and allow you to choose from an existing built-in icon for your own. Unfortunately, there isn’t a lot out there on the interwebs that helps you figure out what the built-in types are. There’s an MSDN article that describes at length a way to create a program that lists all the icons. But I don’t want to write a program to figure them out! Just show them to me! Sheesh :) Thankfully, someone out there felt the same way, and uses a novel hack to get the icons to show up in an outlook toolbar. He then painstakingly took screenshots of them, one group at a time. It isn’t complete though – there are tens of thousands of icons. But it’s good enough. If anyone has an exhaustive list, please let me, and the rest of the add-in community know. Icon Face ID Reference Installing the Add-in It will work with Visual Studio 2008 and Visual Studio 2010. Just unzip the release into your Documents\Visual Studio 20xx\Addins folder. It contains the binary and the Visual Studio “.addin” file. For example, the path to mine is: C:\Users\InfinitiesLoop\Documents\Visual Studio 2010\Addins Conclusion So that’s it! I hope you find it as useful as I have. It’s on GitHub, so if you’re into this kind of thing, please do fork it and improve it! Reference: VSNewFile on GitHub VSNewFile release on GitHub Icon Face ID Reference

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  • Investigation: Can different combinations of components effect Dataflow performance?

    - by jamiet
    Introduction The Dataflow task is one of the core components (if not the core component) of SQL Server Integration Services (SSIS) and often the most misunderstood. This is not surprising, its an incredibly complicated beast and we’re abstracted away from that complexity via some boxes that go yellow red or green and that have some lines drawn between them. Example dataflow In this blog post I intend to look under that facade and get into some of the nuts and bolts of the Dataflow Task by investigating how the decisions we make when building our packages can affect performance. I will do this by comparing the performance of three dataflows that all have the same input, all produce the same output, but which all operate slightly differently by way of having different transformation components. I also want to use this blog post to challenge a common held opinion that I see perpetuated over and over again on the SSIS forum. That is, that people assume adding components to a dataflow will be detrimental to overall performance. Its not surprising that people think this –it is intuitive to think that more components means more work- however this is not a view that I share. I have always been of the opinion that there are many factors affecting dataflow duration and the number of components is actually one of the less important ones; having said that I have never proven that assertion and that is one reason for this investigation. I have actually seen evidence that some people think dataflow duration is simply a function of number of rows and number of components. I’ll happily call that one out as a myth even without any investigation!  The Setup I have a 2GB datafile which is a list of 4731904 (~4.7million) customer records with various attributes against them and it contains 2 columns that I am going to use for categorisation: [YearlyIncome] [BirthDate] The data file is a SSIS raw format file which I chose to use because it is the quickest way of getting data into a dataflow and given that I am testing the transformations, not the source or destination adapters, I want to minimise external influences as much as possible. In the test I will split the customers according to month of birth (12 of those) and whether or not their yearly income is above or below 50000 (2 of those); in other words I will be splitting them into 24 discrete categories and in order to do it I shall be using different combinations of SSIS’ Conditional Split and Derived Column transformation components. The 24 datapaths that occur will each input to a rowcount component, again because this is the least resource intensive means of terminating a datapath. The test is being carried out on a Dell XPS Studio laptop with a quad core (8 logical Procs) Intel Core i7 at 1.73GHz and Samsung SSD hard drive. Its running SQL Server 2008 R2 on Windows 7. The Variables Here are the three combinations of components that I am going to test:     One Conditional Split - A single Conditional Split component CSPL Split by Month of Birth and income category that will use expressions on [YearlyIncome] & [BirthDate] to send each row to one of 24 outputs. This next screenshot displays the expression logic in use: Derived Column & Conditional Split - A Derived Column component DER Income Category that adds a new column [IncomeCategory] which will contain one of two possible text values {“LessThan50000”,”GreaterThan50000”} and uses [YearlyIncome] to determine which value each row should get. A Conditional Split component CSPL Split by Month of Birth and Income Category then uses that new column in conjunction with [BirthDate] to determine which of the same 24 outputs to send each row to. Put more simply, I am separating the Conditional Split of #1 into a Derived Column and a Conditional Split. The next screenshots display the expression logic in use: DER Income Category         CSPL Split by Month of Birth and Income Category       Three Conditional Splits - A Conditional Split component that produces two outputs based on [YearlyIncome], one for each Income Category. Each of those outputs will go to a further Conditional Split that splits the input into 12 outputs, one for each month of birth (identical logic in each). In this case then I am separating the single Conditional Split of #1 into three Conditional Split components. The next screenshots display the expression logic in use: CSPL Split by Income Category         CSPL Split by Month of Birth 1& 2       Each of these combinations will provide an input to one of the 24 rowcount components, just the same as before. For illustration here is a screenshot of the dataflow containing three Conditional Split components: As you can these dataflows have a fair bit of work to do and remember that they’re doing that work for 4.7million rows. I will execute each dataflow 10 times and use the average for comparison. I foresee three possible outcomes: The dataflow containing just one Conditional Split (i.e. #1) will be quicker There is no significant difference between any of them One of the two dataflows containing multiple transformation components will be quicker Regardless of which of those outcomes come to pass we will have learnt something and that makes this an interesting test to carry out. Note that I will be executing the dataflows using dtexec.exe rather than hitting F5 within BIDS. The Results and Analysis The table below shows all of the executions, 10 for each dataflow. It also shows the average for each along with a standard deviation. All durations are in seconds. I’m pasting a screenshot because I frankly can’t be bothered with the faffing about needed to make a presentable HTML table. It is plain to see from the average that the dataflow containing three conditional splits is significantly faster, the other two taking 43% and 52% longer respectively. This seems strange though, right? Why does the dataflow containing the most components outperform the other two by such a big margin? The answer is actually quite logical when you put some thought into it and I’ll explain that below. Before progressing, a side note. The standard deviation for the “Three Conditional Splits” dataflow is orders of magnitude smaller – indicating that performance for this dataflow can be predicted with much greater confidence too. The Explanation I refer you to the screenshot above that shows how CSPL Split by Month of Birth and salary category in the first dataflow is setup. Observe that there is a case for each combination of Month Of Date and Income Category – 24 in total. These expressions get evaluated in the order that they appear and hence if we assume that Month of Date and Income Category are uniformly distributed in the dataset we can deduce that the expected number of expression evaluations for each row is 12.5 i.e. 1 (the minimum) + 24 (the maximum) divided by 2 = 12.5. Now take a look at the screenshots for the second dataflow. We are doing one expression evaluation in DER Income Category and we have the same 24 cases in CSPL Split by Month of Birth and Income Category as we had before, only the expression differs slightly. In this case then we have 1 + 12.5 = 13.5 expected evaluations for each row – that would account for the slightly longer average execution time for this dataflow. Now onto the third dataflow, the quick one. CSPL Split by Income Category does a maximum of 2 expression evaluations thus the expected number of evaluations per row is 1.5. CSPL Split by Month of Birth 1 & CSPL Split by Month of Birth 2 both have less work to do than the previous Conditional Split components because they only have 12 cases to test for thus the expected number of expression evaluations is 6.5 There are two of them so total expected number of expression evaluations for this dataflow is 6.5 + 6.5 + 1.5 = 14.5. 14.5 is still more than 12.5 & 13.5 though so why is the third dataflow so much quicker? Simple, the conditional expressions in the first two dataflows have two boolean predicates to evaluate – one for Income Category and one for Month of Birth; the expressions in the Conditional Split in the third dataflow however only have one predicate thus they are doing a lot less work. To sum up, the difference in execution times can be attributed to the difference between: MONTH(BirthDate) == 1 && YearlyIncome <= 50000 and MONTH(BirthDate) == 1 In the first two dataflows YearlyIncome <= 50000 gets evaluated an average of 12.5 times for every row whereas in the third dataflow it is evaluated once and once only. Multiply those 11.5 extra operations by 4.7million rows and you get a significant amount of extra CPU cycles – that’s where our duration difference comes from. The Wrap-up The obvious point here is that adding new components to a dataflow isn’t necessarily going to make it go any slower, moreover you may be able to achieve significant improvements by splitting logic over multiple components rather than one. Performance tuning is all about reducing the amount of work that needs to be done and that doesn’t necessarily mean use less components, indeed sometimes you may be able to reduce workload in ways that aren’t immediately obvious as I think I have proven here. Of course there are many variables in play here and your mileage will most definitely vary. I encourage you to download the package and see if you get similar results – let me know in the comments. The package contains all three dataflows plus a fourth dataflow that will create the 2GB raw file for you (you will also need the [AdventureWorksDW2008] sample database from which to source the data); simply disable all dataflows except the one you want to test before executing the package and remember, execute using dtexec, not within BIDS. If you want to explore dataflow performance tuning in more detail then here are some links you might want to check out: Inequality joins, Asynchronous transformations and Lookups Destination Adapter Comparison Don’t turn the dataflow into a cursor SSIS Dataflow – Designing for performance (webinar) Any comments? Let me know! @Jamiet

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  • MSCC: Purpose and benefits of Version Control Systems (VCS)

    Unfortunately, there was no monthly meetup during May. Which means that it was even more important and interesting to go forward with a great topic for this month. Earlier this year I already spoke to Nayar Joolfoo about doing a presentation on version control systems (VCS), and he gladly agreed since then. It was just about finding the right date for the action. Furthermore, it was also a great coincidence that Avinash Meetoo announced on social media networks that Knowledge 7 is about to have a new training on "Effective git" - which correlates to a book title Avinash is currently working on - all the best with your approach on this and reach out to our MSCC craftsmen for recessions. Once again a big Thank you to Orange Ebene Accelerator on providing the venue for us, and the MSCC members involved on securing the time slot for our event. Unfortunately, it's kind of tough to get an early confirmation for our meetups these days. I'll keep you posted on that one as there are some interesting and exciting options coming up soon. Okay, let's talk about the meeting and version control systems again. As usual, I'm going to put my first impression of the meetup: "Absolutely great topic, questions and discussions on version control systems, like git or VSO. I was also highly pleased by the number of first timers and female IT geeks. Hopefully, we will be able to keep this trend for future get-togethers." And I really have to emphasise the amount of fresh blood coming to our gathering. Also, during the initial phase it was surprising to see that exactly those first-timers, most of them students at various campuses here on the island, had absolutely no idea about version control systems. More about further down... Reactions of other attendees If I counted correctly, we had a total of 17 attendees this month, and I'd like to give you feedback from some of them: "Inspiring. Helped me understand more about GIT." -- Sean on event comments "Joined the meetup today with literally no idea what is a version control system. I have several reasons why I should be starting to use VCS as from NOW in my projects. Thanks Nayar, Jochen and other participants :)" -- Yudish on event comments "Was present today and I'm very satisfied.I was not aware if there was a such tool like git available. Thanks to those who contributed for this meetup.It was great. Learned a lot from this meetup!!" -- Leonardo on event comments "Seriously, I can see how it’s going to ease my task and help me save time. Gone are the issues with files backups.  And since I’ll be doing my dissertation this year, using Git would help me a lot for my backups and I’m grateful to Nayar for the great explanation." -- Swan-Iyah on MSCC meetup : Version Controls Hopefully, I'll be able to get some other sources - personal blogs preferred - on our meeting. Geeks, thank you so much for those encouraging comments. It's really great to experience that we, all members of the MSCC, are doing the right thing to get more IT information out, and to help each other to improve and evolve in our professional careers. Our agenda of the day Honestly, we had a bumpy start... First, I was battling a little bit with the movable room divider in order to maximize the space. I mean, we had 24 RSVPs and usually there might additional people coming along. Then, for what ever reason, we were facing power outages - actually twice in short periods. Not too good for the projector after all, but hey it went smooth for the rest of the time being. And last but not least... our first speaker Nayar got stuck somewhere on the road. ;-) Anyway, not a real show-stopper and we used the time until Nayar's arrival to introduce ourselves a little bit. It is always important for me to get to know the "newbies" a little bit, and as a result we had lots of students of university - first year, second year and recent graduates - among them. Surprisingly, none of them was ever in contact with version control systems at all. I mean, this is a shocking discovery! Similar to the ability of touch-typing I'd say that being able to use (and master) any kind of version control system is compulsory in any job in the IT industry. Seriously, I'm wondering what is being taught during the classes on the campus. All of them have to work on semester assessments or final projects, even in small teams of 2-4 people. That's the perfect occasion to get started with VCS. Already in this phase, we had great input from more experienced VCS users, like Sean, Avinash and myself. git - a modern approach to VCS - Nayar What a tour! Nayar gave us the full round of git from start to finish, even touching some more advanced techniques. First, he started to explain about the importance of version control systems as an essential tool for software developers, even working alone on a project, and the ability to have a kind of "time machine" that allows you to inspect and revert to a previous version of source code at any time. Then he showed how easy it is to install git on an Ubuntu based system but also mentioned that git is literally available for any operating system, like Windows, Mac OS X and of course other Linux distributions. Next, he showed us how to set the initial configuration values of user name and email address which simplifies the daily usage of the git client while working with your repositories. Then he initialised and added a new repository for some local development of a blogging software. All commands were done using the command line interface (CLI) so that they can be repeated on any system as reference. The syntax and the procedure is always the same, and Nayar clearly mentioned this to the attendees. Now, having a git repository in place it was about time to work on some "important" changes on the blogging software - just for the sake of demonstrating the ease of use and power of git. One interesting question came very early: "How many commands do we have to learn? It looks quite difficult at the moment" - Well, rest assured that during daily development circles you will need less than 10 git commands on a regular base: git add, commit, push, pull, checkout, and merge And Nayar demo'd all of them. Much to the delight of everyone he also showed gitk which is the git repository browser. It's an UI tool to display changes in a repository or a selected set of commits. This includes visualizing the commit graph, showing information related to each commit, and the files in the trees of each revision. Using gitk to display and browse information of a local git repository And last but not least, we took advantage of the internet connectivity and reached out to various online portals offering git hosting for free. Nayar showed us how to push the local repository into a remote system on github. Showing the web-based git browser and history handling, and then also explained and demo'd on how to connect to existing online repositories in order to get access to either your own source code or other people's open source projects. Next to github, we also spoke about bitbucket and gitlab as potential online platforms for your projects. Have a look at the conditions and details about their free service packages and what you can get additionally as a paying customer. Usually, you already get a lot of services for up to five users for free but there might be other important aspects that might have an impact on your decision. Anyways, moving git-based repositories between systems is a piece of cake, and changing online platforms is possible at any stage of your development. Visual Studio Online (VSO) - Jochen Well, Nayar literally covered all elements of working with git during his session, including the use of external online platforms. So, what would be the advantage of talking about Visual Studio Online (VSO)? First of all, VSO is "just another" online platform for hosting and managing git repositories on remote systems, equivalent to github, bitbucket, or any other web site. At the moment (of writing), Microsoft also provides a free package of up to five users / developers on a git repository but there is more in that package. Of course, it is related to software development on the Windows systems and the bonds are tightened towards the use of Visual Studio but out of experience you are absolutely not restricted to that. Connecting a Linux or Mac OS X machine with a git client or an integrated development environment (IDE) like Eclipse or Xcode works as smooth as expected. So, why should one opt in for VSO? Well, one of the main aspects that I would like to mention here is that VSO integrates the Application Life Cycle Methodology (ALM) of Microsoft in their platform. Meaning that you get agile project management with Backlogs, Sprints, Burn-down charts as well as the ability to track tasks, bug reports and work items next to collaborative team chats. It's the whole package of agile development you'll get. And, something I mentioned briefly during the begin of our meeting, VSO gives you the possibility of an automated continuous integrated (CI) process which builds and can run tests of your source code after each commit of changes. Having a proper CI strategy is also part of the Clean Code Developer practices - on Level Green actually -, and not only simplifies your life as a software developer but also reduces the sources of potential errors. Seamless integration and automated deployment between Microsoft Azure Web Sites and git repository But my favourite feature is the seamless continuous deployment to Microsoft Azure. Especially, while working on web projects it's absolutely astounishing that as soon as you commit your chances it just takes a couple of seconds until your modifications are deployed and available on your Azure-hosted web sites. Upcoming Events and networking Due to the adjusted times, everybody was kind of hungry and we didn't follow up on networking or upcoming events - very unfortunate to my opinion and this will have an impact on future planning of our meetups. Because I rather would like to see more conversations during and at the end of our meetings than everyone just packing their laptops, bags and accessories and rush off to grab some food. I was hoping to get some information regarding this year's Code Challenge - supposedly to be organised during July? Maybe someone could leave a comment on that - but I couldn't get any updates. Well, I'll keep digging... In case that you would like to get more into git and how to use it effectively, please check out Knowledge 7's upcoming course on "Effective git". Thanks Avinash for your vital input into today's conversation and I'm looking forward to get a grip on your book title very soon. My resume of the day Do not work in IT without any kind of version control system! Seriously, without a VCS in place you're doing it wrong. It's like driving a car without seat belts attached or riding your bike without safety helmet. You don't do that! End of discussion. ;-) Nowadays, having access to free (as in cost) tools to install on your machine and numerous online platforms to host your source code for free for up to five users it's a no-brainer to get yourself familiar with VCS. Today's sessions gave a good overview on how to start using git and how to connect to various remote services like github or VSO.

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  • [GEEK SCHOOL] Network Security 2: Preventing Disaster with User Account Control

    - by Ciprian Rusen
    In this second lesson in our How-To Geek School about securing the Windows devices in your network, we will talk about User Account Control (UAC). Users encounter this feature each time they need to install desktop applications in Windows, when some applications need administrator permissions in order to work and when they have to change different system settings and files. UAC was introduced in Windows Vista as part of Microsoft’s “Trustworthy Computing” initiative. Basically, UAC is meant to act as a wedge between you and installing applications or making system changes. When you attempt to do either of these actions, UAC will pop up and interrupt you. You may either have to confirm you know what you’re doing, or even enter an administrator password if you don’t have those rights. Some users find UAC annoying and choose to disable it but this very important security feature of Windows (and we strongly caution against doing that). That’s why in this lesson, we will carefully explain what UAC is and everything it does. As you will see, this feature has an important role in keeping Windows safe from all kinds of security problems. In this lesson you will learn which activities may trigger a UAC prompt asking for permissions and how UAC can be set so that it strikes the best balance between usability and security. You will also learn what kind of information you can find in each UAC prompt. Last but not least, you will learn why you should never turn off this feature of Windows. By the time we’re done today, we think you will have a newly found appreciation for UAC, and will be able to find a happy medium between turning it off completely and letting it annoy you to distraction. What is UAC and How Does it Work? UAC or User Account Control is a security feature that helps prevent unauthorized system changes to your Windows computer or device. These changes can be made by users, applications, and sadly, malware (which is the biggest reason why UAC exists in the first place). When an important system change is initiated, Windows displays a UAC prompt asking for your permission to make the change. If you don’t give your approval, the change is not made. In Windows, you will encounter UAC prompts mostly when working with desktop applications that require administrative permissions. For example, in order to install an application, the installer (generally a setup.exe file) asks Windows for administrative permissions. UAC initiates an elevation prompt like the one shown earlier asking you whether it is okay to elevate permissions or not. If you say “Yes”, the installer starts as administrator and it is able to make the necessary system changes in order to install the application correctly. When the installer is closed, its administrator privileges are gone. If you run it again, the UAC prompt is shown again because your previous approval is not remembered. If you say “No”, the installer is not allowed to run and no system changes are made. If a system change is initiated from a user account that is not an administrator, e.g. the Guest account, the UAC prompt will also ask for the administrator password in order to give the necessary permissions. Without this password, the change won’t be made. Which Activities Trigger a UAC Prompt? There are many types of activities that may trigger a UAC prompt: Running a desktop application as an administrator Making changes to settings and files in the Windows and Program Files folders Installing or removing drivers and desktop applications Installing ActiveX controls Changing settings to Windows features like the Windows Firewall, UAC, Windows Update, Windows Defender, and others Adding, modifying, or removing user accounts Configuring Parental Controls in Windows 7 or Family Safety in Windows 8.x Running the Task Scheduler Restoring backed-up system files Viewing or changing the folders and files of another user account Changing the system date and time You will encounter UAC prompts during some or all of these activities, depending on how UAC is set on your Windows device. If this security feature is turned off, any user account or desktop application can make any of these changes without a prompt asking for permissions. In this scenario, the different forms of malware existing on the Internet will also have a higher chance of infecting and taking control of your system. In Windows 8.x operating systems you will never see a UAC prompt when working with apps from the Windows Store. That’s because these apps, by design, are not allowed to modify any system settings or files. You will encounter UAC prompts only when working with desktop programs. What You Can Learn from a UAC Prompt? When you see a UAC prompt on the screen, take time to read the information displayed so that you get a better understanding of what is going on. Each prompt first tells you the name of the program that wants to make system changes to your device, then you can see the verified publisher of that program. Dodgy software tends not to display this information and instead of a real company name, you will see an entry that says “Unknown”. If you have downloaded that program from a less than trustworthy source, then it might be better to select “No” in the UAC prompt. The prompt also shares the origin of the file that’s trying to make these changes. In most cases the file origin is “Hard drive on this computer”. You can learn more by pressing “Show details”. You will see an additional entry named “Program location” where you can see the physical location on your hard drive, for the file that’s trying to perform system changes. Make your choice based on the trust you have in the program you are trying to run and its publisher. If a less-known file from a suspicious location is requesting a UAC prompt, then you should seriously consider pressing “No”. What’s Different About Each UAC Level? Windows 7 and Windows 8.x have four UAC levels: Always notify – when this level is used, you are notified before desktop applications make changes that require administrator permissions or before you or another user account changes Windows settings like the ones mentioned earlier. When the UAC prompt is shown, the desktop is dimmed and you must choose “Yes” or “No” before you can do anything else. This is the most secure and also the most annoying way to set UAC because it triggers the most UAC prompts. Notify me only when programs/apps try to make changes to my computer (default) – Windows uses this as the default for UAC. When this level is used, you are notified before desktop applications make changes that require administrator permissions. If you are making system changes, UAC doesn’t show any prompts and it automatically gives you the necessary permissions for making the changes you desire. When a UAC prompt is shown, the desktop is dimmed and you must choose “Yes” or “No” before you can do anything else. This level is slightly less secure than the previous one because malicious programs can be created for simulating the keystrokes or mouse moves of a user and change system settings for you. If you have a good security solution in place, this scenario should never occur. Notify me only when programs/apps try to make changes to my computer (do not dim my desktop) – this level is different from the previous in in the fact that, when the UAC prompt is shown, the desktop is not dimmed. This decreases the security of your system because different kinds of desktop applications (including malware) might be able to interfere with the UAC prompt and approve changes that you might not want to be performed. Never notify – this level is the equivalent of turning off UAC. When using it, you have no protection against unauthorized system changes. Any desktop application and any user account can make system changes without your permission. How to Configure UAC If you would like to change the UAC level used by Windows, open the Control Panel, then go to “System and Security” and select “Action Center”. On the column on the left you will see an entry that says “Change User Account Control settings”. The “User Account Control Settings” window is now opened. Change the position of the UAC slider to the level you want applied then press “OK”. Depending on how UAC was initially set, you may receive a UAC prompt requiring you to confirm this change. Why You Should Never Turn Off UAC If you want to keep the security of your system at decent levels, you should never turn off UAC. When you disable it, everything and everyone can make system changes without your consent. This makes it easier for all kinds of malware to infect and take control of your system. It doesn’t matter whether you have a security suite or antivirus installed or third-party antivirus, basic common-sense measures like having UAC turned on make a big difference in keeping your devices safe from harm. We have noticed that some users disable UAC prior to setting up their Windows devices and installing third-party software on them. They keep it disabled while installing all the software they will use and enable it when done installing everything, so that they don’t have to deal with so many UAC prompts. Unfortunately this causes problems with some desktop applications. They may fail to work after you enable UAC. This happens because, when UAC is disabled, the virtualization techniques UAC uses for your applications are inactive. This means that certain user settings and files are installed in a different place and when you turn on UAC, applications stop working because they should be placed elsewhere. Therefore, whatever you do, do not turn off UAC completely! Coming up next … In the next lesson you will learn about Windows Defender, what this tool can do in Windows 7 and Windows 8.x, what’s different about it in these operating systems and how it can be used to increase the security of your system.

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  • How I Work: A Cloud Developer's Workstation

    - by BuckWoody
    I've written here a little about how I work during the day, including things like using a stand-up desk (still doing that, by the way). Inspired by a Twitter conversation yesterday, I thought I might explain how I set up my computing environment. First, a couple of important points. I work in Cloud Computing, specifically (but not limited to) Windows Azure. Windows Azure has features to run a Virtual Machine (IaaS), run code without having to control a Virtual Machine (PaaS) and use databases, video streaming, Hadoop and more (a kind of SaaS for tech pros). As such, my designs run the gamut of on-premises, VM's in the Cloud, and software that I write for a platform. I focus on data primarily, meaning that I design a lot of systems that use an RDBMS (like SQL Server or Windows Azure Databases) or a NoSQL approach (MongoDB on Azure or large-scale Key-Value Pairs in Table storage) and even Hadoop and R, and also Cloud Numerics in F#. All that being said, those things inform my choices below. Hardware I have a Lenovo X220 tablet/laptop which I really like a great deal - it's a light, tough, extremely fast system. When I travel, that's the system I take. It has 8GB of RAM, and an SSD drive. I sometimes use that to develop or work at a client's site, on the road, or in the living room when I'm not in my home office. My main system is a GateWay DX430017 - I've maxed it out on RAM, and I have two 1TB drives in it. It's not only my workstation for work; I leave it on all the time and it streams our videos, music and books. I have about 3400 e-books, and I've just started using Calibre to stream the library. I run Windows 8 on it so I can set up Hyper-V images, since Windows Azure allows me to move regular Hyper-V disks back and forth to the Cloud. That's where all my "servers" are, when I have to use an IaaS approach. The reason I use a desktop-style system rather than a laptop only approach is that a good part of my job is setting up architectures to solve really big, complex problems. That means I have to simulate entire networks on-premises, along with the Hybrid Cloud approach I use a lot. I need a lot of disk space and memory for that, and I use two huge monitors on my stand-up desk. I could probably use 10 monitors if I had the room for them. Also, since it's our home system as well, I leave it on all the time and it doesn't travel.   Software For the software for my systems, it's important to keep in mind that I not only write code, but I design databases, teach, present, and create Linux and other environments. Windows 8 - While the jury is out for me on the new interface, the context-sensitive search, integrated everything, and speed is just hands-down the right choice. I've evaluated a server OS, Linux, even an Apple, but I just am not as efficient on those as I am with Windows 8. Visual Studio Ultimate - I develop primarily in .NET (C# and F# mostly) and I use the Team Foundation Server in the cloud, and I'm asked to do everything from UI to Services, so I need everything. Windows Azure SDK, Windows Azure Training Kit - I need the first to set up my Azure PaaS coding, and the second has all the info I need for PaaS, IaaS and SaaS. This is primarily how I get paid. :) SQL Server Developer Edition - While I might install Oracle, MySQL and Postgres on my VM's, the "outside" environment is SQL Server for an RDBMS. I install the Developer Edition because it has the same features as Enterprise Edition, and comes with all the client tools and documentation. Microsoft Office -  Even if I didn't work here, this is what I would use. I've just grown too accustomed to doing business this way to change, so my advice is always "use what works", and this does. The parts I use are: OneNote (and a Math Add-in) - I do almost everything - and I mean everything in OneNote. I can code, do high-end math, present, design, collaborate and more. All my notebooks are on my Skydrive. I can use them from any system, anywhere. If you take the time to learn this program, you'll be hooked. Excel with PowerPivot - Don't make that face. Excel is the world's database, and every Data Scientist I know - even the ones where I teach at the University of Washington - know it, use it, and love it.  Outlook - Primary communications, CRM and contact tool. I have all of my social media hooked up to it, so when I get an e-mail from you, I see everything, see all the history we've had on e-mail, find you on a map and more. Lync - I was fine with LiveMeeting, although it has it's moments. For me, the Lync client is tres-awesome. I use this throughout my day, present on it, stay in contact with colleagues and the folks on the dev team (who wish I didn't have it) and more.  PowerPoint - Once again, don't make that face. Whenever I see someone complaining about PowerPoint, I have 100% of the time found they don't know how to use it. If you suck at presenting or creating content, don't blame PowerPoint. Works great on my machine. :) Zoomit - Magnifier - On Windows 7 (and 8 as well) there's a built-in magnifier, but I install Zoomit out of habit. It enlarges the screen. If you don't use one of these tools (or their equivalent on some other OS) then you're presenting/teaching wrong, and you should stop presenting/teaching until you get them and learn how to show people what you can see on your tiny, tiny monitor. :) Cygwin - Unix for Windows. OK, that's not true, but it's mostly that. I grew up on mainframes and Unix (IBM and HP, thank you) and I can't imagine life without  sed, awk, grep, vim, and bash. I also tend to take a lot of the "Science" and "Development" and "Database" packages in it as well. PuTTY - Speaking of Unix, when I need to connect to my Linux VM's in Windows Azure, I want to do it securely. This is the tool for that. Notepad++ - Somewhere between torturing myself in vim and luxuriating in OneNote is Notepad++. Everyone has a favorite text editor; this one is mine. Too many features to name, and it's free. Browsers - I install Chrome, Firefox and of course IE. I know it's in vogue to rant on IE, but I tend to think for myself a great deal, and I've had few (none) problems with it. The others I have for the haterz that make sites that won't run in IE. Visio - I've used a lot of design packages, but none have the extreme meta-data edit capabilities of Visio. I don't use this all the time - it can be rather heavy, but what it does it does really well. I also present this way when I'm not using PowerPoint. Yup, I just bring up Visio and diagram away as I'm chatting with clients. Depending on what we're covering, this can be the right tool for that. Tweetdeck - The AIR one, not that new disaster they came out with. I live on social media, since you, dear readers, are my cube-mates. When I get tired of you all, I close Tweetdeck. When I need help or someone needs help from me, or if I want to see a picture of a cat while I'm coding, I bring it up. It's up most all day and night. Windows Media Player - I listen to Trance or Classical when I code, and I find music managers overbearing and extra. I just use what comes in the box, and it works great for me. R - F# and Cloud Numerics now allows me to load in R libraries (yay!) and I use this for statistical work on big data loads. Microsoft Math - One of the most amazing, free, rich, amazing, awesome, amazing calculators out there. I get the 64-bit version for quick math conversions, plots and formula-checks. Python - I know, right? Who knew that the scientific community loved Python so much. But they do. I use 2.7; not as much runs with 3+. I also use IronPython in Visual Studio, or I edit in Notepad++ Camstudio recorder - Windows PSR - In much of my training, and all of my teaching at the UW, I need to show a process on a screen. Camstudio records screen and voice, and it's free. If I need to make static training, I use the Windows PSR tool that's built right in. It's ostensibly for problem duplication, but I use it to record for training.   OK - your turn. Post a link to your blog entry below, and tell me how you set your system up.  

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  • Installing SharePoint 2010 and PowerPivot for SharePoint on Windows 7

    - by smisner
    Many people like me want (or need) to do their business intelligence development work on a laptop. As someone who frequently speaks at various events or teaches classes on all subjects related to the Microsoft business intelligence stack, I need a way to run multiple server products on my laptop with reasonable performance. Once upon a time, that requirement meant only that I had to load the current version of SQL Server and the client tools of choice. In today's post, I'll review my latest experience with trying to make the newly released Microsoft BI products work with a Windows 7 operating system. The entrance of Microsoft Office SharePoint Server 2007 into the BI stack complicated matters and I started using Virtual Server to establish a "suitable" environment. As part of the team that delivered a lot of education as part of the Yukon pre-launch activities (that would be SQL Server 2005 for the uninitiated), I was working with four - yes, four - virtual servers. That was a pretty brutal workload for a 2GB laptop, which worked if I was very, very careful. It could also be a finicky and unreliable configuration as I learned to my dismay at one TechEd session several years ago when I had to reboot a very carefully cached set of servers just minutes before my session started. Although it worked, it came back to life very, very slowly much to the displeasure of the audience. They couldn't possibly have been less pleased than me. At that moment, I resolved to get the beefiest environment I could afford and consolidate to a single virtual server. Enter the 4GB 64-bit laptop to preserve my sanity and my livelihood. Likewise, for SQL Server 2008, I managed to keep everything within a single virtual server and I could function reasonably well with this approach. Now we have SQL Server 2008 R2 plus Office SharePoint Server 2010. That means a 64-bit operating system. Period. That means no more Virtual Server. That means I must use Hyper-V or another alternative. I've heard alternatives exist, but my few dabbles in this area did not yield positive results. It might have been just me having issues rather than any failure of those technologies to adequately support the requirements. My first run at working with the new BI stack configuration was to set up a 64-bit 4GB laptop with a dual-boot to run Windows Server 2008 R2 with Hyper-V. However, I was generally not happy with running Windows Server 2008 R2 on my laptop. For one, I couldn't put it into sleep mode, which is helpful if I want to prepare for a presentation beforehand and then walk to the podium without the need to hold my laptop in its open state along the way (my strategy at the TechEd session long, long ago). Secondly, it was finicky with projectors. I had issues from time to time and while I always eventually got it to work, I didn't appreciate those nerve-wracking moments wondering whether this would be the time that it wouldn't work. Somewhere along the way, I learned that it was possible to load SharePoint 2010 in a Windows 7 which piqued my interest. I had just acquired a new laptop running Windows 7 64-bit, and thought surely running the BI stack natively on my laptop must be better than running Hyper-V. (I have not tried booting to Hyper-V VHD yet, but that's on my list of things to try so the jury of one is still out on this approach.) Recently, I had to build up a server with the RTM versions of SQL Server 2008 R2 and Sharepoint Server 2010 and decided to follow suit on my Windows 7 Ultimate 64-bit laptop. The process is slightly different, but I'm happy to report that it IS possible, although I had some fits and starts along the way. DISCLAIMER: These products are NOT intended to be run in production mode on the Windows 7 operating system. The configuration described in this post is strictly for development or learning purposes and not supported by Microsoft. If you have trouble, you will NOT get help from them. I might be able to help, but I provide no guarantees of my ability or availablity to help. I won't provide the step-by-step instructions in this post as there are other resources that provide these details, but I will provide an overview of my approach, point you to the relevant resources, describe some of the problems I encountered, and explain how I addressed those problems to achieve my desired goal. Because my goal was not simply to set up SharePoint Server 2010 on my laptop, but specifically PowerPivot for SharePoint, I started out by referring to the installation instructions at the PowerPiovt-Info site, but mainly to confirm that I was performing steps in the proper sequence. I didn't perform the steps in Part 1 because those steps are applicable only to a server operating system which I am not running on my laptop. Then, the instructions in Part 2, won't work exactly as written for the same reason. Instead, I followed the instructions on MSDN, Setting Up the Development Environment for SharePoint 2010 on Windows Vista, Windows 7, and Windows Server 2008. In general, I found the following differences in installation steps from the steps at PowerPivot-Info: You must copy the SharePoint installation media to the local drive so that you can edit the config.xml to allow installation on a Windows client. You also have to manually install the prerequisites. The instructions provides links to each item that you must manually install and provides a command-line instruction to execute which enables required Windows features. I will digress for a moment to save you some grief in the sequence of steps to perform. I discovered later that a missing step in the MSDN instructions is to install the November CTP Reporting Services add-in for SharePoint. When I went to test my SharePoint site (I believe I tested after I had a successful PowerPivot installation), I ran into the following error: Could not load file or assembly 'RSSharePointSoapProxy, Version=10.0.0.0, Culture=neutral, PublicKeyToken=89845dcd8080cc91' or one of its dependencies. The system cannot find the file specified. I was rather surprised that Reporting Services was required. Then I found an article by Alan le Marquand, Working Together: SQL Server 2008 R2 Reporting Services Integration in SharePoint 2010,that instructed readers to install the November add-in. My first reaction was, "Really?!?" But I confirmed it in another TechNet article on hardware and software requirements for SharePoint Server 2010. It doesn't refer explicitly to the November CTP but following the link took me there. (Interestingly, I retested today and there's no longer any reference to the November CTP. Here's the link to download the latest and greatest Reporting Services Add-in for SharePoint Technologies 2010.) You don't need to download the add-in anymore if you're doing a regular server-based installation of SharePoint because it installs as part of the prerequisites automatically. When it was time to start the installation of SharePoint, I deviated from the MSDN instructions and from the PowerPivot-Info instructions: On the Choose the installation you want page of the installation wizard, I chose Server Farm. On the Server Type page, I chose Complete. At the end of the installation, I did not run the configuration wizard. Returning to the PowerPivot-Info instructions, I tried to follow the instructions in Part 3 which describe installing SQL Server 2008 R2 with the PowerPivot option. These instructions tell you to choose the New Server option on the Setup Role page where you add PowerPivot for SharePoint. However, I ran into problems with this approach and got installation errors at the end. It wasn't until much later as I was investigating an error that I encountered Dave Wickert's post that installing PowerPivot for SharePoint on Windows 7 is unsupported. Uh oh. But he did want to hear about it if anyone succeeded, so I decided to take the plunge. Perseverance paid off, and I can happily inform Dave that it does work so far. I haven't tested absolutely everything with PowerPivot for SharePoint but have successfully deployed a workbook and viewed the PowerPivot Management Dashboard. I have not yet tested the data refresh feature, but I have installed. Continue reading to see how I accomplished my objective. I unintalled SQL Server 2008 R2 and started again. I had different problems which I don't recollect now. However, I uninstalled again and approached installation from a different angle and my next attempt succeeded. The downside of this approach is that you must do all of the things yourself that are done automatically when you install PowerPivot as a new server. Here are the steps that I followed: Install SQL Server 2008 R2 to get a database engine instance installed. Run the SharePoint configuration wizard to set up the SharePoint databases. In Central Administration, create a Web application using classic mode authentication as per a TechNet article on PowerPivot Authentication and Authorization. Then I followed the steps I found at How to: Install PowerPivot for SharePoint on an Existing SharePoint Server. Especially important to note - you must launch setup by using Run as administrator. I did not have to manually deploy the PowerPivot solution as the instructions specify, but it's good to know about this step because it tells you where to look in Central Administration to confirm a successful deployment. I did spot some incorrect steps in the instructions (at the time of this writing) in How To: Configure Stored Credentials for PowerPivot Data Refresh. Specifically, in the section entitled Step 1: Create a target application and set the credentials, both steps 10 and 12 are incorrect. They tell you to provide an actual Windows user name and password on the page where you are simply defining the prompts for your application in the Secure Store Service. To add the Windows user name and password that you want to associate with the application - after you have successfully created the target application - you select the target application and then click Set credentials in the ribbon. Lastly, I followed the instructions at How to: Install Office Data Connectivity Components on a PowerPivot server. However, I have yet to test this in my current environment. I did have several stops and starts throughout this process and edited those out to spare you from reading non-essential information. I believe the explanation I have provided here accurately reflect the steps I followed to produce a working configuration. If you follow these steps and get a different result, please let me know so that together we can work through the issue and correct these instructions. I'm sure there are many other folks in the Microsoft BI community that will appreciate the ability to set up the BI stack in a Windows 7 environment for development or learning purposes. Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • Much Ado About Nothing: Stub Objects

    - by user9154181
    The Solaris 11 link-editor (ld) contains support for a new type of object that we call a stub object. A stub object is a shared object, built entirely from mapfiles, that supplies the same linking interface as the real object, while containing no code or data. Stub objects cannot be executed — the runtime linker will kill any process that attempts to load one. However, you can link to a stub object as a dependency, allowing the stub to act as a proxy for the real version of the object. You may well wonder if there is a point to producing an object that contains nothing but linking interface. As it turns out, stub objects are very useful for building large bodies of code such as Solaris. In the last year, we've had considerable success in applying them to one of our oldest and thorniest build problems. In this discussion, I will describe how we came to invent these objects, and how we apply them to building Solaris. This posting explains where the idea for stub objects came from, and details our long and twisty journey from hallway idea to standard link-editor feature. I expect that these details are mainly of interest to those who work on Solaris and its makefiles, those who have done so in the past, and those who work with other similar bodies of code. A subsequent posting will omit the history and background details, and instead discuss how to build and use stub objects. If you are mainly interested in what stub objects are, and don't care about the underlying software war stories, I encourage you to skip ahead. The Long Road To Stubs This all started for me with an email discussion in May of 2008, regarding a change request that was filed in 2002, entitled: 4631488 lib/Makefile is too patient: .WAITs should be reduced This CR encapsulates a number of cronic issues with Solaris builds: We build Solaris with a parallel make (dmake) that tries to build as much of the code base in parallel as possible. There is a lot of code to build, and we've long made use of parallelized builds to get the job done quicker. This is even more important in today's world of massively multicore hardware. Solaris contains a large number of executables and shared objects. Executables depend on shared objects, and shared objects can depend on each other. Before you can build an object, you need to ensure that the objects it needs have been built. This implies a need for serialization, which is in direct opposition to the desire to build everying in parallel. To accurately build objects in the right order requires an accurate set of make rules defining the things that depend on each other. This sounds simple, but the reality is quite complex. In practice, having programmers explicitly specify these dependencies is a losing strategy: It's really hard to get right. It's really easy to get it wrong and never know it because things build anyway. Even if you get it right, it won't stay that way, because dependencies between objects can change over time, and make cannot help you detect such drifing. You won't know that you got it wrong until the builds break. That can be a long time after the change that triggered the breakage happened, making it hard to connect the cause and the effect. Usually this happens just before a release, when the pressure is on, its hard to think calmly, and there is no time for deep fixes. As a poor compromise, the libraries in core Solaris were built using a set of grossly incomplete hand written rules, supplemented with a number of dmake .WAIT directives used to group the libraries into sets of non-interacting groups that can be built in parallel because we think they don't depend on each other. From time to time, someone will suggest that we could analyze the built objects themselves to determine their dependencies and then generate make rules based on those relationships. This is possible, but but there are complications that limit the usefulness of that approach: To analyze an object, you have to build it first. This is a classic chicken and egg scenario. You could analyze the results of a previous build, but then you're not necessarily going to get accurate rules for the current code. It should be possible to build the code without having a built workspace available. The analysis will take time, and remember that we're constantly trying to make builds faster, not slower. By definition, such an approach will always be approximate, and therefore only incremantally more accurate than the hand written rules described above. The hand written rules are fast and cheap, while this idea is slow and complex, so we stayed with the hand written approach. Solaris was built that way, essentially forever, because these are genuinely difficult problems that had no easy answer. The makefiles were full of build races in which the right outcomes happened reliably for years until a new machine or a change in build server workload upset the accidental balance of things. After figuring out what had happened, you'd mutter "How did that ever work?", add another incomplete and soon to be inaccurate make dependency rule to the system, and move on. This was not a satisfying solution, as we tend to be perfectionists in the Solaris group, but we didn't have a better answer. It worked well enough, approximately. And so it went for years. We needed a different approach — a new idea to cut the Gordian Knot. In that discussion from May 2008, my fellow linker-alien Rod Evans had the initial spark that lead us to a game changing series of realizations: The link-editor is used to link objects together, but it only uses the ELF metadata in the object, consisting of symbol tables, ELF versioning sections, and similar data. Notably, it does not look at, or understand, the machine code that makes an object useful at runtime. If you had an object that only contained the ELF metadata for a dependency, but not the code or data, the link-editor would find it equally useful for linking, and would never know the difference. Call it a stub object. In the core Solaris OS, we require all objects to be built with a link-editor mapfile that describes all of its publically available functions and data. Could we build a stub object using the mapfile for the real object? It ought to be very fast to build stub objects, as there are no input objects to process. Unlike the real object, stub objects would not actually require any dependencies, and so, all of the stubs for the entire system could be built in parallel. When building the real objects, one could link against the stub objects instead of the real dependencies. This means that all the real objects can be built built in parallel too, without any serialization. We could replace a system that requires perfect makefile rules with a system that requires no ordering rules whatsoever. The results would be considerably more robust. We immediately realized that this idea had potential, but also that there were many details to sort out, lots of work to do, and that perhaps it wouldn't really pan out. As is often the case, it would be necessary to do the work and see how it turned out. Following that conversation, I set about trying to build a stub object. We determined that a faithful stub has to do the following: Present the same set of global symbols, with the same ELF versioning, as the real object. Functions are simple — it suffices to have a symbol of the right type, possibly, but not necessarily, referencing a null function in its text segment. Copy relocations make data more complicated to stub. The possibility of a copy relocation means that when you create a stub, the data symbols must have the actual size of the real data. Any error in this will go uncaught at link time, and will cause tragic failures at runtime that are very hard to diagnose. For reasons too obscure to go into here, involving tentative symbols, it is also important that the data reside in bss, or not, matching its placement in the real object. If the real object has more than one symbol pointing at the same data item, we call these aliased symbols. All data symbols in the stub object must exhibit the same aliasing as the real object. We imagined the stub library feature working as follows: A command line option to ld tells it to produce a stub rather than a real object. In this mode, only mapfiles are examined, and any object or shared libraries on the command line are are ignored. The extra information needed (function or data, size, and bss details) would be added to the mapfile. When building the real object instead of the stub, the extra information for building stubs would be validated against the resulting object to ensure that they match. In exploring these ideas, I immediately run headfirst into the reality of the original mapfile syntax, a subject that I would later write about as The Problem(s) With Solaris SVR4 Link-Editor Mapfiles. The idea of extending that poor language was a non-starter. Until a better mapfile syntax became available, which seemed unlikely in 2008, the solution could not involve extentions to the mapfile syntax. Instead, we cooked up the idea (hack) of augmenting mapfiles with stylized comments that would carry the necessary information. A typical definition might look like: # DATA(i386) __iob 0x3c0 # DATA(amd64,sparcv9) __iob 0xa00 # DATA(sparc) __iob 0x140 iob; A further problem then became clear: If we can't extend the mapfile syntax, then there's no good way to extend ld with an option to produce stub objects, and to validate them against the real objects. The idea of having ld read comments in a mapfile and parse them for content is an unacceptable hack. The entire point of comments is that they are strictly for the human reader, and explicitly ignored by the tool. Taking all of these speed bumps into account, I made a new plan: A perl script reads the mapfiles, generates some small C glue code to produce empty functions and data definitions, compiles and links the stub object from the generated glue code, and then deletes the generated glue code. Another perl script used after both objects have been built, to compare the real and stub objects, using data from elfdump, and validate that they present the same linking interface. By June 2008, I had written the above, and generated a stub object for libc. It was a useful prototype process to go through, and it allowed me to explore the ideas at a deep level. Ultimately though, the result was unsatisfactory as a basis for real product. There were so many issues: The use of stylized comments were fine for a prototype, but not close to professional enough for shipping product. The idea of having to document and support it was a large concern. The ideal solution for stub objects really does involve having the link-editor accept the same arguments used to build the real object, augmented with a single extra command line option. Any other solution, such as our prototype script, will require makefiles to be modified in deeper ways to support building stubs, and so, will raise barriers to converting existing code. A validation script that rederives what the linker knew when it built an object will always be at a disadvantage relative to the actual linker that did the work. A stub object should be identifyable as such. In the prototype, there was no tag or other metadata that would let you know that they weren't real objects. Being able to identify a stub object in this way means that the file command can tell you what it is, and that the runtime linker can refuse to try and run a program that loads one. At that point, we needed to apply this prototype to building Solaris. As you might imagine, the task of modifying all the makefiles in the core Solaris code base in order to do this is a massive task, and not something you'd enter into lightly. The quality of the prototype just wasn't good enough to justify that sort of time commitment, so I tabled the project, putting it on my list of long term things to think about, and moved on to other work. It would sit there for a couple of years. Semi-coincidentally, one of the projects I tacked after that was to create a new mapfile syntax for the Solaris link-editor. We had wanted to do something about the old mapfile syntax for many years. Others before me had done some paper designs, and a great deal of thought had already gone into the features it should, and should not have, but for various reasons things had never moved beyond the idea stage. When I joined Sun in late 2005, I got involved in reviewing those things and thinking about the problem. Now in 2008, fresh from relearning for the Nth time why the old mapfile syntax was a huge impediment to linker progress, it seemed like the right time to tackle the mapfile issue. Paving the way for proper stub object support was not the driving force behind that effort, but I certainly had them in mind as I moved forward. The new mapfile syntax, which we call version 2, integrated into Nevada build snv_135 in in February 2010: 6916788 ld version 2 mapfile syntax PSARC/2009/688 Human readable and extensible ld mapfile syntax In order to prove that the new mapfile syntax was adequate for general purpose use, I had also done an overhaul of the ON consolidation to convert all mapfiles to use the new syntax, and put checks in place that would ensure that no use of the old syntax would creep back in. That work went back into snv_144 in June 2010: 6916796 OSnet mapfiles should use version 2 link-editor syntax That was a big putback, modifying 517 files, adding 18 new files, and removing 110 old ones. I would have done this putback anyway, as the work was already done, and the benefits of human readable syntax are obvious. However, among the justifications listed in CR 6916796 was this We anticipate adding additional features to the new mapfile language that will be applicable to ON, and which will require all sharable object mapfiles to use the new syntax. I never explained what those additional features were, and no one asked. It was premature to say so, but this was a reference to stub objects. By that point, I had already put together a working prototype link-editor with the necessary support for stub objects. I was pleased to find that building stubs was indeed very fast. On my desktop system (Ultra 24), an amd64 stub for libc can can be built in a fraction of a second: % ptime ld -64 -z stub -o stubs/libc.so.1 -G -hlibc.so.1 \ -ztext -zdefs -Bdirect ... real 0.019708910 user 0.010101680 sys 0.008528431 In order to go from prototype to integrated link-editor feature, I knew that I would need to prove that stub objects were valuable. And to do that, I knew that I'd have to switch the Solaris ON consolidation to use stub objects and evaluate the outcome. And in order to do that experiment, ON would first need to be converted to version 2 mapfiles. Sub-mission accomplished. Normally when you design a new feature, you can devise reasonably small tests to show it works, and then deploy it incrementally, letting it prove its value as it goes. The entire point of stub objects however was to demonstrate that they could be successfully applied to an extremely large and complex code base, and specifically to solve the Solaris build issues detailed above. There was no way to finesse the matter — in order to move ahead, I would have to successfully use stub objects to build the entire ON consolidation and demonstrate their value. In software, the need to boil the ocean can often be a warning sign that things are trending in the wrong direction. Conversely, sometimes progress demands that you build something large and new all at once. A big win, or a big loss — sometimes all you can do is try it and see what happens. And so, I spent some time staring at ON makefiles trying to get a handle on how things work, and how they'd have to change. It's a big and messy world, full of complex interactions, unspecified dependencies, special cases, and knowledge of arcane makefile features... ...and so, I backed away, put it down for a few months and did other work... ...until the fall, when I felt like it was time to stop thinking and pondering (some would say stalling) and get on with it. Without stubs, the following gives a simplified high level view of how Solaris is built: An initially empty directory known as the proto, and referenced via the ROOT makefile macro is established to receive the files that make up the Solaris distribution. A top level setup rule creates the proto area, and performs operations needed to initialize the workspace so that the main build operations can be launched, such as copying needed header files into the proto area. Parallel builds are launched to build the kernel (usr/src/uts), libraries (usr/src/lib), and commands. The install makefile target builds each item and delivers a copy to the proto area. All libraries and executables link against the objects previously installed in the proto, implying the need to synchronize the order in which things are built. Subsequent passes run lint, and do packaging. Given this structure, the additions to use stub objects are: A new second proto area is established, known as the stub proto and referenced via the STUBROOT makefile macro. The stub proto has the same structure as the real proto, but is used to hold stub objects. All files in the real proto are delivered as part of the Solaris product. In contrast, the stub proto is used to build the product, and then thrown away. A new target is added to library Makefiles called stub. This rule builds the stub objects. The ld command is designed so that you can build a stub object using the same ld command line you'd use to build the real object, with the addition of a single -z stub option. This means that the makefile rules for building the stub objects are very similar to those used to build the real objects, and many existing makefile definitions can be shared between them. A new target is added to the Makefiles called stubinstall which delivers the stub objects built by the stub rule into the stub proto. These rules reuse much of existing plumbing used by the existing install rule. The setup rule runs stubinstall over the entire lib subtree as part of its initialization. All libraries and executables link against the objects in the stub proto rather than the main proto, and can therefore be built in parallel without any synchronization. There was no small way to try this that would yield meaningful results. I would have to take a leap of faith and edit approximately 1850 makefiles and 300 mapfiles first, trusting that it would all work out. Once the editing was done, I'd type make and see what happened. This took about 6 weeks to do, and there were many dark days when I'd question the entire project, or struggle to understand some of the many twisted and complex situations I'd uncover in the makefiles. I even found a couple of new issues that required changes to the new stub object related code I'd added to ld. With a substantial amount of encouragement and help from some key people in the Solaris group, I eventually got the editing done and stub objects for the entire workspace built. I found that my desktop system could build all the stub objects in the workspace in roughly a minute. This was great news, as it meant that use of the feature is effectively free — no one was likely to notice or care about the cost of building them. After another week of typing make, fixing whatever failed, and doing it again, I succeeded in getting a complete build! The next step was to remove all of the make rules and .WAIT statements dedicated to controlling the order in which libraries under usr/src/lib are built. This came together pretty quickly, and after a few more speed bumps, I had a workspace that built cleanly and looked like something you might actually be able to integrate someday. This was a significant milestone, but there was still much left to do. I turned to doing full nightly builds. Every type of build (open, closed, OpenSolaris, export, domestic) had to be tried. Each type failed in a new and unique way, requiring some thinking and rework. As things came together, I became aware of things that could have been done better, simpler, or cleaner, and those things also required some rethinking, the seeking of wisdom from others, and some rework. After another couple of weeks, it was in close to final form. My focus turned towards the end game and integration. This was a huge workspace, and needed to go back soon, before changes in the gate would made merging increasingly difficult. At this point, I knew that the stub objects had greatly simplified the makefile logic and uncovered a number of race conditions, some of which had been there for years. I assumed that the builds were faster too, so I did some builds intended to quantify the speedup in build time that resulted from this approach. It had never occurred to me that there might not be one. And so, I was very surprised to find that the wall clock build times for a stock ON workspace were essentially identical to the times for my stub library enabled version! This is why it is important to always measure, and not just to assume. One can tell from first principles, based on all those removed dependency rules in the library makefile, that the stub object version of ON gives dmake considerably more opportunities to overlap library construction. Some hypothesis were proposed, and shot down: Could we have disabled dmakes parallel feature? No, a quick check showed things being build in parallel. It was suggested that we might be I/O bound, and so, the threads would be mostly idle. That's a plausible explanation, but system stats didn't really support it. Plus, the timing between the stub and non-stub cases were just too suspiciously identical. Are our machines already handling as much parallelism as they are capable of, and unable to exploit these additional opportunities? Once again, we didn't see the evidence to back this up. Eventually, a more plausible and obvious reason emerged: We build the libraries and commands (usr/src/lib, usr/src/cmd) in parallel with the kernel (usr/src/uts). The kernel is the long leg in that race, and so, wall clock measurements of build time are essentially showing how long it takes to build uts. Although it would have been nice to post a huge speedup immediately, we can take solace in knowing that stub objects simplify the makefiles and reduce the possibility of race conditions. The next step in reducing build time should be to find ways to reduce or overlap the uts part of the builds. When that leg of the build becomes shorter, then the increased parallelism in the libs and commands will pay additional dividends. Until then, we'll just have to settle for simpler and more robust. And so, I integrated the link-editor support for creating stub objects into snv_153 (November 2010) with 6993877 ld should produce stub objects PSARC/2010/397 ELF Stub Objects followed by the work to convert the ON consolidation in snv_161 (February 2011) with 7009826 OSnet should use stub objects 4631488 lib/Makefile is too patient: .WAITs should be reduced This was a huge putback, with 2108 modified files, 8 new files, and 2 removed files. Due to the size, I was allowed a window after snv_160 closed in which to do the putback. It went pretty smoothly for something this big, a few more preexisting race conditions would be discovered and addressed over the next few weeks, and things have been quiet since then. Conclusions and Looking Forward Solaris has been built with stub objects since February. The fact that developers no longer specify the order in which libraries are built has been a big success, and we've eliminated an entire class of build error. That's not to say that there are no build races left in the ON makefiles, but we've taken a substantial bite out of the problem while generally simplifying and improving things. The introduction of a stub proto area has also opened some interesting new possibilities for other build improvements. As this article has become quite long, and as those uses do not involve stub objects, I will defer that discussion to a future article.

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  • Issue 15: SVP Focus

    - by rituchhibber
         SVP FOCUS FOCUS -- Chris Baker SVP Oracle Worldwide ISV-OEM-Java Sales Chris Baker is the Global Head of ISV/OEM Sales responsible for working with ISV/OEM partners to maximise Oracle's business through those partners, whilst maximising those partners’ business to their end users. Chris works with partners, customers, innovators, investors and employees to develop innovative business solutions using Oracle products, services and skills. RESOURCES -- Oracle PartnerNetwork (OPN) OPN Solutions Catalog Oracle Exastack Program Oracle Exastack Optimized Oracle Cloud Computing Oracle Engineered Systems Oracle and Java SUBSCRIBE FEEDBACK PREVIOUS ISSUES "By taking part in marketing activities, our partners accelerate their sales cycles." -- Firstly, could you please explain Oracle's current strategy for ISV partners, globally and in EMEA? Oracle customers use independent software vendor (ISV) applications to run their businesses. They use them to generate revenue and to fulfil obligations to their own customers. Our strategy is very straight-forward. We want all of our ISV partners and OEMs to concentrate on the things that they do the best—building applications to meet the unique industry and functional requirements of their customer. We want to ensure that we deliver a best-in-class application platform so ISVs are free to concentrate their effort on their application functionality and user experience We invest over four billion dollars in research and development every year, and we want our ISVs to benefit from all of that investment in operating systems, virtualisation, databases, middleware, engineered systems, and other hardware. By doing this, we help them to reduce their costs, gain more consistency and agility for quicker implementations, and also rapidly differentiate themselves from other application vendors. It's all about simplification because we believe that around 25 to 30 percent of the development costs incurred by many ISVs are caused by customising infrastructure and have nothing to do with their applications. Our strategy is to enable our ISV partners to standardise their application platform using engineered architecture, so they can write once to the Oracle stack and deploy seamlessly in the cloud, on-premise, or in hybrid deployments. It's really important that architecture is the same in order to keep cost and time overheads at a minimum, so we provide standardisation and an environment that enables our ISVs to concentrate on the core business that makes them the most money and brings them success. How do you believe this strategy is helping the ISVs to work hand-in-hand with Oracle to ensure that end customers get the industry-leading solutions that they need? We work with our ISVs not just to help them be successful, but also to help them market themselves. We have something called the 'Oracle Exastack Ready Program', which enables ISVs to publicise themselves as 'Ready' to run the core software platforms that run on Oracle's engineered systems including Exadata and Exalogic. So, for example, they can become 'Database Ready' which means that they use the latest version of Oracle Database and therefore can run their application without modification on Exadata or the Oracle Database Appliance. Alternatively, they can become WebLogic Ready, Oracle Linux Ready and Oracle Solaris Ready which means they run on the latest release and therefore can run their application, with no new porting work, on Oracle Exalogic. Those 'Ready' logos are important in helping ISVs advertise to their customers that they are using the latest technologies which have been fully tested. We now also have Exadata Ready and Exalogic Ready programmes which allow ISVs to promote the certification of their applications on these platforms. This highlights these partners to Oracle customers as having solutions that run fluently on the Oracle Exadata Database Machine, the Oracle Exalogic Elastic Cloud or one of our other engineered systems. This makes it easy for customers to identify solutions and provides ISVs with an avenue to connect with Oracle customers who are rapidly adopting engineered systems. We have also taken this programme to the next level in the shape of 'Oracle Exastack Optimized' for partners whose applications run best on the Oracle stack and have invested the time to fully optimise application performance. We ensure that Exastack Optimized partner status is promoted and supported by press releases, and we help our ISVs go to market and differentiate themselves through the use of our technology and the standardisation it delivers. To date we have had several hundred organisations successfully work through our Exastack Optimized programme. How does Oracle's strategy of offering pre-integrated open platform software and hardware allow ISVs to bring their products to market more quickly? One of the problems for many ISVs is that they have to think very carefully about the technology on which their solutions will be deployed, particularly in the cloud or hosted environments. They have to think hard about how they secure these environments, whether the concern is, for example, middleware, identity management, or securing personal data. If they don't use the technology that we build-in to our products to help them to fulfil these roles, they then have to build it themselves. This takes time, requires testing, and must be maintained. By taking advantage of our technology, partners will now know that they have a standard platform. They will know that they can confidently talk about implementation being the same every time they do it. Very large ISV applications could once take a year or two to be implemented at an on-premise environment. But it wasn't just the configuration of the application that took the time, it was actually the infrastructure - the different hardware configurations, operating systems and configurations of databases and middleware. Now we strongly believe that it's all about standardisation and repeatability. It's about making sure that our partners can do it once and are then able to roll it out many different times using standard componentry. What actions would you recommend for existing ISV partners that are looking to do more business with Oracle and its customer base, not only to maximise benefits, but also to maximise partner relationships? My team, around the world and in the EMEA region, is available and ready to talk to any of our ISVs and to explore the possibilities together. We run programmes like 'Excite' and 'Insight' to help us to understand how we can help ISVs with architecture and widen their environments. But we also want to work with, and look at, new opportunities - for example, the Machine-to-Machine (M2M) market or 'The Internet of Things'. Over the next few years, many millions, indeed billions of devices will be collecting massive amounts of data and communicating it back to the central systems where ISVs will be running their applications. The only way that our partners will be able to provide a single vendor 'end-to-end' solution is to use Oracle integrated systems at the back end and Java on the 'smart' devices collecting the data—a complete solution from device to data centre. So there are huge opportunities to work closely with our ISVs, using Oracle's complete M2M platform, to provide the infrastructure that enables them to extract maximum value from the data collected. If any partners don't know where to start or who to contact, then they can contact me directly at [email protected] or indeed any of our teams across the EMEA region. We want to work with ISVs to help them to be as successful as they possibly can through simplification and speed to market, and we also want all of the top ISVs in the world based on Oracle. What opportunities are immediately opened to new ISV partners joining the OPN? As you know OPN is very, very important. New members will discover a huge amount of content that instantly becomes accessible to them. They can access a wealth of no-cost training and enablement materials to build their expertise in Oracle technology. They can download Oracle software and use it for development projects. They can help themselves become more competent by becoming part of a true community and uncovering new opportunities by working with Oracle and their peers in the Oracle Partner Network. As well as publishing massive amounts of information on OPN, we also hold our global Oracle OpenWorld event, at which partners play a huge role. This takes place at the end of September and the beginning of October in San Francisco. Attending ISV partners have an unrivalled opportunity to contribute to elements such as the OpenWorld / OPN Exchange, at which they can talk to other partners and really begin thinking about how they can move their businesses on and play key roles in a very large ecosystem which revolves around technology and standardisation. Finally, are there any other messages that you would like to share with the Oracle ISV community? The crucial message that I always like to reinforce is architecture, architecture and architecture! The key opportunities that ISVs have today revolve around standardising their architectures so that they can confidently think: "I will I be able to do exactly the same thing whenever a customer is looking to deploy on-premise, hosted or in the cloud". The right architecture is critical to being competitive and to really start changing the game. We want to help our ISV partners to do just that; to establish standard architecture and to seize the opportunities it opens up for them. New market opportunities like M2M are enormous - just look at how many devices are all around you right now. We can help our partners to interface with these devices more effectively while thinking about their entire ecosystem, rather than just the piece that they have traditionally focused upon. With standardised architecture, we can help people dramatically improve their speed, reach, agility and delivery of enhanced customer satisfaction and value all the way from the Java side to their centralised systems. All Oracle ISV partners must take advantage of these opportunities, which is why Oracle will continue to invest in and support them. Oracle OpenWorld 2010 Whether you attended Oracle OpenWorld 2009 or not, don't forget to save the date now for Oracle OpenWorld 2010. The event will be held a little earlier next year, from 19th-23rd September, so please don't miss out. With thousands of sessions and hundreds of exhibits and demos already lined up, there's no better place to learn how to optimise your existing systems, get an inside line on upcoming technology breakthroughs, and meet with your partner peers, Oracle strategists and even the developers responsible for the products and services that help you get better results for your end customers. Register Now for Oracle OpenWorld 2010! Perhaps you are interested in learning more about Oracle OpenWorld 2010, but don't wish to register at this time? Great! Please just enter your contact information here and we will contact you at a later date. How to Exhibit at Oracle OpenWorld 2010 Sponsorship Opportunities at Oracle OpenWorld 2010 Advertising Opportunities at Oracle OpenWorld 2010 -- Back to the welcome page

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  • how this scaling down for css code is worked?

    - by harris
    this is a code for scaling down for css. i was wondering, how this worked. please someone explain to me part by part. thank you very much. /* ======================================================================== / / Copyright (C) 2000 - 2009 ND-Tech. Co., Ltd. / / All Rights Reserved. / / ======================================================================== / / Project : ScaleDown Created : 31-AUG-2009 / / File : main.c Contact : [email protected] / / ======================================================================== / / You are free to use or modify this code to the following restrictions: / / Acknowledge ND Tech. Co. Ltd. / / Or, put "Parts of code by ND Tech. Co., Ltd." / / Or, leave this header as it is. / / in somewhere in your code. / / ======================================================================== */ include "vm3224k.h" define CE0CTL *(volatile int *)(0x01800008) define CE2CTL *(volatile int *)(0x01800010) define SDCTL *(volatile int *)(0x01800018) define LED *(volatile short *)(0x90080000) // Definitions for async access(change as you wish) define WSU (2<<28) // Write Setup : 0-15 define WST (8<<22) // Write Strobe: 0-63 define WHD (2<<20) // Write Hold : 0-3 define RSU (2<<16) // Read Setup : 0-15 define TA (3<<14) // Turn Around : 0-3 define RST (8<<8) // Read Strobe : 0-63 define RHD (2<<0) // Read Hold : 0-3 define MTYPE (2<<4) /* EDMA Registers */ define PaRAM_OPT 0 // Options define PaRAM_SRC 1 // Source Address define PaRAM_CNT 2 // Frame count, Element count define PaRAM_DST 3 // Destination Address define PaRAM_IDX 4 // Frame index, Element index define PaRAM_RDL 5 // Element count reload, Link address define EDMA_CIPR *(volatile int *)0x01A0FFE4 // EDMA Channel interrupt pending low register define EDMA_CIER *(volatile int *)0x01A0FFE8 // EDMA Channel interrupt enable low register define EDMA_CCER *(volatile int *)0x01A0FFEC // EDMA Channel chain enable register define EDMA_ER *(volatile int *)0x01A0FFF0 // EDMA Event low register define EDMA_EER *(volatile int *)0x01A0FFF4 // EDMA Event enable low register define EDMA_ECR *(volatile int *)0x01A0FFF8 // EDMA Event clear low register define EDMA_ESR *(volatile int *)0x01A0FFFC // EDMA Event set low register define PRI (2<<29) // 1:High priority, 2:Low priority define ESIZE (1<<27) // 0:32bit, 1:16bit, 2:8bit, 3:reserved define DS2 (0<<26) // 1:2-Dimensional define SUM (0<<24) // 0:no update, 1:increment, 2:decrement, 3:by index define DD2 (0<<23) // 1:2-Dimensional define DUM (0<<21) // 0:no update, 1:increment, 2:decrement, 3:by index define TCINT (1<<20) // 0:disable, 1:enable define TCC (8<<16) // 4 bit code define LINK (0<<1) // 0:disable, 1:enable define FS (1<<0) // 0:element, 1:frame define OptionField_0 (PRI|ESIZE|DS2|SUM|DD2|DUM|TCINT|TCC|LINK|FS) define DD2_1 (1<<23) // 1:2-Dimensional define DUM_1 (1<<21) // 0:no update, 1:increment, 2:decrement, 3:by index define TCC_1 (9<<16) // 4 bit code define OptionField_1 (PRI|ESIZE|DS2|SUM|DD2_1|DUM_1|TCINT|TCC_1|LINK|FS) define TCC_2 (10<<16)// 4 bit code define OptionField_2 (PRI|ESIZE|DS2|SUM|DD2|DUM|TCINT|TCC_2|LINK|FS) define DS2_3 (1<<26) // 1:2-Dimensional define SUM_3 (1<<24) // 0:no update, 1:increment, 2:decrement, 3:by index define TCC_3 (11<<16)// 4 bit code define OptionField_3 (PRI|ESIZE|DS2_3|SUM_3|DD2|DUM|TCINT|TCC_3|LINK|FS) pragma DATA_SECTION ( lcd,".sdram" ) pragma DATA_SECTION ( cam,".sdram" ) pragma DATA_SECTION ( rgb,".sdram" ) pragma DATA_SECTION ( u,".sdram" ) extern cregister volatile unsigned int IER; extern cregister volatile unsigned int CSR; short camcode = 0x08000; short lcdcode = 0x00000; short lcd[2][240][320]; short cam[2][240][320]; short rgb[64][32][32]; short bufsel; int *Cevent,*Levent,*CLink,flag=1; unsigned char v[240][160],out_y[120][160]; unsigned char y[240][320],out_u[120][80]; unsigned char u[240][160],out_v[120][80]; void PLL6713() { int i; // CPU Clock Input : 50MHz *(volatile int *)(0x01b7c100) = *(volatile int *)(0x01b7c100) & 0xfffffffe; for(i=0;i<4;i++); *(volatile int *)(0x01b7c100) = *(volatile int *)(0x01b7c100) | 0x08; *(volatile int *)(0x01b7c114) = 0x08001; // 50MHz/2 = 25MHz *(volatile int *)(0x01b7c110) = 0x0c; // 25MHz * 12 = 300MHz *(volatile int *)(0x01b7c118) = 0x08000; // SYSCLK1 = 300MHz/1 = 300MHz *(volatile int *)(0x01b7c11c) = 0x08001; // SYSCLK2 = 300MHz/2 = 150MHz // Peripheral Clock *(volatile int *)(0x01b7c120) = 0x08003; // SYSCLK3 = 300MHz/4 = 75MHz // SDRAM Clock for(i=0;i<4;i++); *(volatile int *)(0x01b7c100) = *(volatile int *)(0x01b7c100) & 0xfffffff7; for(i=0;i<4;i++); *(volatile int *)(0x01b7c100) = *(volatile int *)(0x01b7c100) | 0x01; } unsigned short ybr_565(short y,short u,short v) { int r,g,b; b = y + 1772*(u-128)/1000; if (b<0) b=0; if (b>255) b=255; g = y - (344*(u-128) + 714*(v-128))/1000; if (g<0) g=0; if (g>255) g=255; r = y + 1402*(v-128)/1000; if (r<0) r=0; if (r>255) r=255; return ((r&0x0f8)<<8)|((g&0x0fc)<<3)|((b&0x0f8)>>3); } void yuyv2yuv(char *yuyv,char *y,char *u,char *v) { int i,j,dy,dy1,dy2,s; for (j=s=dy=dy1=dy2=0;j<240;j++) { for (i=0;i<320;i+=2) { u[dy1++] = yuyv[s++]; y[dy++] = yuyv[s++]; v[dy2++] = yuyv[s++]; y[dy++] = yuyv[s++]; } } } interrupt void c_int06(void) { if(EDMA_CIPR&0x800){ EDMA_CIPR = 0xffff; bufsel=(++bufsel&0x01); Cevent[PaRAM_DST] = (int)cam[(bufsel+1)&0x01]; Levent[PaRAM_SRC] = (int)lcd[(bufsel+1)&0x01]; EDMA_ESR = 0x80; flag=1; } } void main() { int i,j,k,y0,y1,v0,u0; bufsel = 0; CSR &= (~0x1); PLL6713(); // Initialize C6713 PLL CE0CTL = 0xffffbf33;// SDRAM Space CE2CTL = (WSU|WST|WHD|RSU|RST|RHD|MTYPE); SDCTL = 0x57115000; vm3224init(); // Initialize vm3224k2 vm3224rate(1); // Set frame rate vm3224bl(15); // Set backlight VM3224CNTL = VM3224CNTL&0xffff | 0x2; // vm3224 interrupt enable for (k=0;k<64;k++) // Create RGB565 lookup table for (i=0;i<32;i++) for (j=0;j<32;j++) rgb[k][i][j] = ybr_565(k<<2,i<<3,j<<3); Cevent = (int *)(0x01a00000 + 24 * 7); Cevent[PaRAM_OPT] = OptionField_0; Cevent[PaRAM_SRC] = (int)&camcode; Cevent[PaRAM_CNT] = 1; Cevent[PaRAM_DST] = (int)&VM3224ADDH; Cevent = (int *)(0x01a00000 + 24 * 8); Cevent[PaRAM_OPT] = OptionField_1; Cevent[PaRAM_SRC] = (int)&VM3224DATA; Cevent[PaRAM_CNT] = (239<<16)|320; Cevent[PaRAM_DST] = (int)cam[bufsel]; Cevent[PaRAM_IDX] = 0; Levent = (int *)(0x01a00000 + 24 * 9); Levent[PaRAM_OPT] = OptionField_2; Levent[PaRAM_SRC] = (int)&lcdcode; Levent[PaRAM_CNT] = 1; Levent[PaRAM_DST] = (int)&VM3224ADDH; Levent = (int *)(0x01a00000 + 24 * 10); Levent[PaRAM_OPT] = OptionField_3; Levent[PaRAM_SRC] = (int)lcd[bufsel]; Levent[PaRAM_CNT] = (239<<16)|320; Levent[PaRAM_DST] = (int)&VM3224DATA; Levent[PaRAM_IDX] = 0; IER = IER | (1<<6)|3; CSR = CSR | 0x1; EDMA_CCER = (1<<8)|(1<<9)|(1<<10); EDMA_CIER = (1<<11); EDMA_CIPR = 0xffff; EDMA_ESR = 0x80; while (1) { if(flag) { // LED = 0; yuyv2yuv((char *)cam[bufsel],(char *)y,(char *)u,(char *)v); for(j=0;j<240;j++) for(i=0;i<320;i++) lcd[bufsel][j][i]=0; for(j=0;j<240;j+=2) for(i=0;i<320;i+=2) out_y[j>>1][i>>1]=(y[j][i]+y[j][i+1]+y[j+1][i]+y[j+1][i+1])>>2; for(j=0;j<240;j+=2) for(i=0;i<160;i+=2) { out_u[j>>1][i>>1]=(u[j][i]+u[j][i+1]+u[j+1][i]+u[j+1][i+1])>>2; out_v[j>>1][i>>1]=(v[j][i]+v[j][i+1]+v[j+1][i]+v[j+1][i+1])>>2; } for (j=0;j<120;j++) for (i=0;i<160;i+=2) { y0 = out_y[j][i]>>2; u0 = out_u[j][i>>1]>>3; v0 = out_v[j][i>>1]>>3; y1 = out_y[j][i+1]>>2; lcd[bufsel][j+60][i+80]=rgb[y0][u0][v0]; lcd[bufsel][j+60][i+81]=rgb[y1][u0][v0]; } flag=0; // LED = 1; } } }

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  • SATA errors reported during boot: exception Emask 0x40 SAct 0x0 SErr 0x80800 action 0x0

    - by digby280
    I have noticed some error during the Linux boot. They seem to continue to occur after the boot adding lines to the log every few seconds. Once booted this normally does not appear to be causing any problems. However, around 1 in 10 boots results in a kernel panic and the computer has on two or three occasions suddenly rebooted after being powered on for a number of hours. I presume the cause of the reboot is a kernel panic as well. I am running Ubuntu 11.10 and I have had Ubuntu installed on the computer for around a year. I have googled around and not found anything useful. I have provided the kernel log lines and the output of smartctl. Can anyone explain exactly what these errors mean, or better still how to resolve them? Apr 2 16:51:27 dell580 kernel: [ 19.831140] EXT4-fs (sdb2): re-mounted. Opts: errors=remount-ro,user_xattr,commit=0 Apr 2 16:51:27 dell580 kernel: [ 19.934194] tg3 0000:03:00.0: eth0: Link is down Apr 2 16:51:28 dell580 kernel: [ 20.929468] tg3 0000:03:00.0: eth0: Link is up at 100 Mbps, full duplex Apr 2 16:51:28 dell580 kernel: [ 20.929471] tg3 0000:03:00.0: eth0: Flow control is on for TX and on for RX Apr 2 16:51:28 dell580 kernel: [ 20.929727] ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready Apr 2 16:51:29 dell580 kernel: [ 21.609381] EXT4-fs (sdb2): re-mounted. Opts: errors=remount-ro,user_xattr,commit=0 Apr 2 16:51:29 dell580 kernel: [ 21.616515] ata2.01: exception Emask 0x40 SAct 0x0 SErr 0x80800 action 0x0 Apr 2 16:51:29 dell580 kernel: [ 21.616519] ata2.01: SError: { HostInt 10B8B } Apr 2 16:51:29 dell580 kernel: [ 21.616525] ata2.00: hard resetting link Apr 2 16:51:29 dell580 kernel: [ 21.934036] ata2.01: hard resetting link Apr 2 16:51:29 dell580 kernel: [ 22.408890] ata2.00: SATA link up 1.5 Gbps (SStatus 113 SControl 300) Apr 2 16:51:29 dell580 kernel: [ 22.408907] ata2.01: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 2 16:51:29 dell580 kernel: [ 22.440934] ata2.00: configured for UDMA/100 Apr 2 16:51:29 dell580 kernel: [ 22.449040] ata2.01: configured for UDMA/133 Apr 2 16:51:29 dell580 kernel: [ 22.449818] ata2: EH complete Apr 2 16:51:33 dell580 kernel: [ 26.122664] ata2.01: exception Emask 0x40 SAct 0x0 SErr 0x80800 action 0x0 Apr 2 16:51:33 dell580 kernel: [ 26.122670] ata2.01: SError: { HostInt 10B8B } Apr 2 16:51:33 dell580 kernel: [ 26.122677] ata2.00: hard resetting link Apr 2 16:51:33 dell580 kernel: [ 26.442684] ata2.01: hard resetting link Apr 2 16:51:34 dell580 kernel: [ 26.925545] ata2.00: SATA link up 1.5 Gbps (SStatus 113 SControl 300) Apr 2 16:51:34 dell580 kernel: [ 26.925561] ata2.01: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 2 16:51:34 dell580 kernel: [ 26.961542] ata2.00: configured for UDMA/100 Apr 2 16:51:34 dell580 kernel: [ 26.969616] ata2.01: configured for UDMA/133 Apr 2 16:51:34 dell580 kernel: [ 26.970400] ata2: EH complete Apr 2 16:51:35 dell580 kernel: [ 28.111180] ata2.01: exception Emask 0x40 SAct 0x0 SErr 0x80800 action 0x0 Apr 2 16:51:35 dell580 kernel: [ 28.111184] ata2.01: SError: { HostInt 10B8B } Apr 2 16:51:35 dell580 kernel: [ 28.111191] ata2.00: hard resetting link Apr 2 16:51:35 dell580 kernel: [ 28.429674] ata2.01: hard resetting link Apr 2 16:51:36 dell580 kernel: [ 28.904557] ata2.00: SATA link up 1.5 Gbps (SStatus 113 SControl 300) Apr 2 16:51:36 dell580 kernel: [ 28.904572] ata2.01: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 2 16:51:36 dell580 kernel: [ 28.936609] ata2.00: configured for UDMA/100 Apr 2 16:51:36 dell580 kernel: [ 28.944692] ata2.01: configured for UDMA/133 Apr 2 16:51:36 dell580 kernel: [ 28.945464] ata2: EH complete Apr 2 16:51:38 dell580 kernel: [ 31.581756] eth0: no IPv6 routers present Apr 2 16:51:38 dell580 kernel: [ 32.103066] ata2.01: exception Emask 0x40 SAct 0x0 SErr 0x80800 action 0x0 Apr 2 16:51:38 dell580 kernel: [ 32.103074] ata2.01: SError: { HostInt 10B8B } Apr 2 16:51:38 dell580 kernel: [ 32.103085] ata2.00: hard resetting link Apr 2 16:51:38 dell580 kernel: [ 32.419669] ata2.01: hard resetting link Apr 2 16:51:39 dell580 kernel: [ 32.894518] ata2.00: SATA link up 1.5 Gbps (SStatus 113 SControl 300) Apr 2 16:51:39 dell580 kernel: [ 32.894533] ata2.01: SATA link up 3.0 Gbps (SStatus 123 SControl 300) Apr 2 16:51:39 dell580 kernel: [ 32.926536] ata2.00: configured for UDMA/100 Apr 2 16:51:39 dell580 kernel: [ 32.934715] ata2.01: configured for UDMA/133 Apr 2 16:51:39 dell580 kernel: [ 32.935578] ata2: EH complete Here's the output of smartctl for the drive. smartctl 5.41 2011-06-09 r3365 [x86_64-linux-3.0.0-17-generic] (local build) Copyright (C) 2002-11 by Bruce Allen, http://smartmontools.sourceforge.net === START OF INFORMATION SECTION === Model Family: SAMSUNG SpinPoint F1 DT Device Model: SAMSUNG HD103UJ Serial Number: S13PJ90QC19706 LU WWN Device Id: 5 0000f0 00b1c7960 Firmware Version: 1AA01113 User Capacity: 1,000,204,886,016 bytes [1.00 TB] Sector Size: 512 bytes logical/physical Device is: In smartctl database [for details use: -P show] ATA Version is: 8 ATA Standard is: ATA-8-ACS revision 3b Local Time is: Mon Apr 2 17:13:48 2012 BST SMART support is: Available - device has SMART capability. SMART support is: Enabled === START OF READ SMART DATA SECTION === SMART overall-health self-assessment test result: PASSED General SMART Values: Offline data collection status: (0x00) Offline data collection activity was never started. Auto Offline Data Collection: Disabled. Self-test execution status: ( 41) The self-test routine was interrupted by the host with a hard or soft reset. Total time to complete Offline data collection: (11772) seconds. Offline data collection capabilities: (0x7b) SMART execute Offline immediate. Auto Offline data collection on/off support. Suspend Offline collection upon new command. Offline surface scan supported. Self-test supported. Conveyance Self-test supported. Selective Self-test supported. SMART capabilities: (0x0003) Saves SMART data before entering power-saving mode. Supports SMART auto save timer. Error logging capability: (0x01) Error logging supported. General Purpose Logging supported. Short self-test routine recommended polling time: ( 2) minutes. Extended self-test routine recommended polling time: ( 197) minutes. Conveyance self-test routine recommended polling time: ( 21) minutes. SCT capabilities: (0x003f) SCT Status supported. SCT Error Recovery Control supported. SCT Feature Control supported. SCT Data Table supported. SMART Attributes Data Structure revision number: 16 Vendor Specific SMART Attributes with Thresholds: ID# ATTRIBUTE_NAME FLAG VALUE WORST THRESH TYPE UPDATED WHEN_FAILED RAW_VALUE 1 Raw_Read_Error_Rate 0x000f 100 100 051 Pre-fail Always - 0 3 Spin_Up_Time 0x0007 076 076 011 Pre-fail Always - 7940 4 Start_Stop_Count 0x0032 099 099 000 Old_age Always - 521 5 Reallocated_Sector_Ct 0x0033 100 100 010 Pre-fail Always - 0 7 Seek_Error_Rate 0x000f 253 253 051 Pre-fail Always - 0 8 Seek_Time_Performance 0x0025 100 100 015 Pre-fail Offline - 0 9 Power_On_Hours 0x0032 100 100 000 Old_age Always - 642 10 Spin_Retry_Count 0x0033 100 100 051 Pre-fail Always - 0 11 Calibration_Retry_Count 0x0012 100 100 000 Old_age Always - 0 12 Power_Cycle_Count 0x0032 100 100 000 Old_age Always - 482 13 Read_Soft_Error_Rate 0x000e 100 100 000 Old_age Always - 0 183 Runtime_Bad_Block 0x0032 100 100 000 Old_age Always - 759 184 End-to-End_Error 0x0033 100 100 000 Pre-fail Always - 0 187 Reported_Uncorrect 0x0032 100 100 000 Old_age Always - 0 188 Command_Timeout 0x0032 100 100 000 Old_age Always - 0 190 Airflow_Temperature_Cel 0x0022 073 069 000 Old_age Always - 27 (Min/Max 16/27) 194 Temperature_Celsius 0x0022 073 067 000 Old_age Always - 27 (Min/Max 16/28) 195 Hardware_ECC_Recovered 0x001a 100 100 000 Old_age Always - 320028 196 Reallocated_Event_Count 0x0032 100 100 000 Old_age Always - 0 197 Current_Pending_Sector 0x0012 100 100 000 Old_age Always - 0 198 Offline_Uncorrectable 0x0030 100 100 000 Old_age Offline - 0 199 UDMA_CRC_Error_Count 0x003e 099 099 000 Old_age Always - 1494 200 Multi_Zone_Error_Rate 0x000a 100 100 000 Old_age Always - 0 201 Soft_Read_Error_Rate 0x000a 253 253 000 Old_age Always - 0 SMART Error Log Version: 1 ATA Error Count: 211 (device log contains only the most recent five errors) CR = Command Register [HEX] FR = Features Register [HEX] SC = Sector Count Register [HEX] SN = Sector Number Register [HEX] CL = Cylinder Low Register [HEX] CH = Cylinder High Register [HEX] DH = Device/Head Register [HEX] DC = Device Command Register [HEX] ER = Error register [HEX] ST = Status register [HEX] Powered_Up_Time is measured from power on, and printed as DDd+hh:mm:SS.sss where DD=days, hh=hours, mm=minutes, SS=sec, and sss=millisec. It "wraps" after 49.710 days. Error 211 occurred at disk power-on lifetime: 0 hours (0 days + 0 hours) When the command that caused the error occurred, the device was active or idle. After command completion occurred, registers were: ER ST SC SN CL CH DH -- -- -- -- -- -- -- 84 51 0f 31 63 8f e1 Error: ICRC, ABRT 15 sectors at LBA = 0x018f6331 = 26174257 Commands leading to the command that caused the error were: CR FR SC SN CL CH DH DC Powered_Up_Time Command/Feature_Name -- -- -- -- -- -- -- -- ---------------- -------------------- c8 00 00 40 62 8f e1 08 00:01:00.460 READ DMA c8 00 20 00 7c 30 e0 08 00:01:00.450 READ DMA c8 00 00 10 49 8f e1 08 00:01:00.440 READ DMA c8 00 e0 20 d0 30 e0 08 00:01:00.420 READ DMA c8 00 00 c0 59 90 e1 08 00:01:00.400 READ DMA Error 210 occurred at disk power-on lifetime: 0 hours (0 days + 0 hours) When the command that caused the error occurred, the device was active or idle. After command completion occurred, registers were: ER ST SC SN CL CH DH -- -- -- -- -- -- -- 84 51 cf e9 cf 66 e0 Error: ICRC, ABRT 207 sectors at LBA = 0x0066cfe9 = 6737897 Commands leading to the command that caused the error were: CR FR SC SN CL CH DH DC Powered_Up_Time Command/Feature_Name -- -- -- -- -- -- -- -- ---------------- -------------------- c8 00 00 b8 cf 66 e0 08 00:08:29.780 READ DMA c8 00 60 60 c9 18 e0 08 00:08:29.770 READ DMA c8 00 40 20 c9 18 e0 08 00:08:29.770 READ DMA c8 00 20 00 c9 18 e0 08 00:08:29.760 READ DMA c8 00 20 98 cf 66 e0 08 00:08:29.750 READ DMA Error 209 occurred at disk power-on lifetime: 0 hours (0 days + 0 hours) When the command that caused the error occurred, the device was active or idle. After command completion occurred, registers were: ER ST SC SN CL CH DH -- -- -- -- -- -- -- 84 51 2f d1 74 e0 e0 Error: ICRC, ABRT 47 sectors at LBA = 0x00e074d1 = 14709969 Commands leading to the command that caused the error were: CR FR SC SN CL CH DH DC Powered_Up_Time Command/Feature_Name -- -- -- -- -- -- -- -- ---------------- -------------------- c8 00 00 00 74 e0 e0 08 00:00:30.940 READ DMA c8 00 20 18 36 de e0 08 00:00:30.930 READ DMA c8 00 08 48 f1 dd e0 08 00:00:30.930 READ DMA c8 00 08 a8 f0 dd e0 08 00:00:30.930 READ DMA c8 00 08 90 f0 dd e0 08 00:00:30.930 READ DMA Error 208 occurred at disk power-on lifetime: 0 hours (0 days + 0 hours) When the command that caused the error occurred, the device was active or idle. After command completion occurred, registers were: ER ST SC SN CL CH DH -- -- -- -- -- -- -- 84 51 7f 21 88 9d e0 Error: ICRC, ABRT 127 sectors at LBA = 0x009d8821 = 10324001 Commands leading to the command that caused the error were: CR FR SC SN CL CH DH DC Powered_Up_Time Command/Feature_Name -- -- -- -- -- -- -- -- ---------------- -------------------- c8 00 a0 00 88 9d e0 08 00:00:27.610 READ DMA c8 00 58 a8 e7 9c e0 08 00:00:27.610 READ DMA c8 00 00 28 e6 9c e0 08 00:00:27.610 READ DMA c8 00 00 e0 e4 9c e0 08 00:00:27.610 READ DMA c8 00 00 90 e0 9c e0 08 00:00:27.600 READ DMA Error 207 occurred at disk power-on lifetime: 0 hours (0 days + 0 hours) When the command that caused the error occurred, the device was active or idle. After command completion occurred, registers were: ER ST SC SN CL CH DH -- -- -- -- -- -- -- 04 51 26 6a 6a c3 e0 Error: ABRT at LBA = 0x00c36a6a = 12806762 Commands leading to the command that caused the error were: CR FR SC SN CL CH DH DC Powered_Up_Time Command/Feature_Name -- -- -- -- -- -- -- -- ---------------- -------------------- ca 00 00 90 69 c3 e0 08 00:29:39.350 WRITE DMA ca 00 40 90 68 c3 e0 08 00:29:39.350 WRITE DMA ca 00 40 50 65 c3 e0 08 00:29:39.350 WRITE DMA ca 00 40 d0 64 c3 e0 08 00:29:39.350 WRITE DMA ca 00 40 90 63 c3 e0 08 00:29:39.350 WRITE DMA SMART Self-test log structure revision number 1 Num Test_Description Status Remaining LifeTime(hours) LBA_of_first_error # 1 Short offline Interrupted (host reset) 90% 638 - # 2 Short offline Interrupted (host reset) 90% 638 - # 3 Extended offline Interrupted (host reset) 90% 638 - # 4 Short offline Interrupted (host reset) 90% 638 - # 5 Extended offline Interrupted (host reset) 90% 638 - SMART Selective self-test log data structure revision number 1 SPAN MIN_LBA MAX_LBA CURRENT_TEST_STATUS 1 0 0 Not_testing 2 0 0 Not_testing 3 0 0 Not_testing 4 0 0 Not_testing 5 0 0 Not_testing Selective self-test flags (0x0): After scanning selected spans, do NOT read-scan remainder of disk. If Selective self-test is pending on power-up, resume after 0 minute delay.

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  • Metro: Dynamically Switching Templates with a WinJS ListView

    - by Stephen.Walther
    Imagine that you want to display a list of products using the WinJS ListView control. Imagine, furthermore, that you want to use different templates to display different products. In particular, when a product is on sale, you want to display the product using a special “On Sale” template. In this blog entry, I explain how you can switch templates dynamically when displaying items with a ListView control. In other words, you learn how to use more than one template when displaying items with a ListView control. Creating the Data Source Let’s start by creating the data source for the ListView. Nothing special here – our data source is a list of products. Two of the products, Oranges and Apples, are on sale. (function () { "use strict"; var products = new WinJS.Binding.List([ { name: "Milk", price: 2.44 }, { name: "Oranges", price: 1.99, onSale: true }, { name: "Wine", price: 8.55 }, { name: "Apples", price: 2.44, onSale: true }, { name: "Steak", price: 1.99 }, { name: "Eggs", price: 2.44 }, { name: "Mushrooms", price: 1.99 }, { name: "Yogurt", price: 2.44 }, { name: "Soup", price: 1.99 }, { name: "Cereal", price: 2.44 }, { name: "Pepsi", price: 1.99 } ]); WinJS.Namespace.define("ListViewDemos", { products: products }); })(); The file above is saved with the name products.js and referenced by the default.html page described below. Declaring the Templates and ListView Control Next, we need to declare the ListView control and the two Template controls which we will use to display template items. The markup below appears in the default.html file: <!-- Templates --> <div id="productItemTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> <div id="productOnSaleTemplate" data-win-control="WinJS.Binding.Template"> <div class="product onSale"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> </div> <!-- ListView --> <div id="productsListView" data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource: ListViewDemos.products.dataSource, layout: { type: WinJS.UI.ListLayout } }"> </div> In the markup above, two Template controls are declared. The first template is used when rendering a normal product and the second template is used when rendering a product which is on sale. The second template, unlike the first template, includes the text “(On Sale!)”. The ListView control is bound to the data source which we created in the previous section. The ListView itemDataSource property is set to the value ListViewDemos.products.dataSource. Notice that we do not set the ListView itemTemplate property. We set this property in the default.js file. Switching Between Templates All of the magic happens in the default.js file. The default.js file contains the JavaScript code used to switch templates dynamically. Here’s the entire contents of the default.js file: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { WinJS.UI.processAll().then(function () { var productsListView = document.getElementById("productsListView"); productsListView.winControl.itemTemplate = itemTemplateFunction; });; } }; function itemTemplateFunction(itemPromise) { return itemPromise.then(function (item) { // Select either normal product template or on sale template var itemTemplate = document.getElementById("productItemTemplate"); if (item.data.onSale) { itemTemplate = document.getElementById("productOnSaleTemplate"); }; // Render selected template to DIV container var container = document.createElement("div"); itemTemplate.winControl.render(item.data, container); return container; }); } app.start(); })(); In the code above, a function is assigned to the ListView itemTemplate property with the following line of code: productsListView.winControl.itemTemplate = itemTemplateFunction;   The itemTemplateFunction returns a DOM element which is used for the template item. Depending on the value of the product onSale property, the DOM element is generated from either the productItemTemplate or the productOnSaleTemplate template. Using Binding Converters instead of Multiple Templates In the previous sections, I explained how you can use different templates to render normal products and on sale products. There is an alternative approach to displaying different markup for normal products and on sale products. Instead of creating two templates, you can create a single template which contains separate DIV elements for a normal product and an on sale product. The following default.html file contains a single item template and a ListView control bound to the template. <!-- Template --> <div id="productItemTemplate" data-win-control="WinJS.Binding.Template"> <div class="product" data-win-bind="style.display: onSale ListViewDemos.displayNormalProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> <div class="product onSale" data-win-bind="style.display: onSale ListViewDemos.displayOnSaleProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> </div> <!-- ListView --> <div id="productsListView" data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource: ListViewDemos.products.dataSource, itemTemplate: select('#productItemTemplate'), layout: { type: WinJS.UI.ListLayout } }"> </div> The first DIV element is used to render a normal product: <div class="product" data-win-bind="style.display: onSale ListViewDemos.displayNormalProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> The second DIV element is used to render an “on sale” product: <div class="product onSale" data-win-bind="style.display: onSale ListViewDemos.displayOnSaleProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> Notice that both templates include a data-win-bind attribute. These data-win-bind attributes are used to show the “normal” template when a product is not on sale and show the “on sale” template when a product is on sale. These attributes set the Cascading Style Sheet display attribute to either “none” or “block”. The data-win-bind attributes take advantage of binding converters. The binding converters are defined in the default.js file: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { WinJS.UI.processAll(); } }; WinJS.Namespace.define("ListViewDemos", { displayNormalProduct: WinJS.Binding.converter(function (onSale) { return onSale ? "none" : "block"; }), displayOnSaleProduct: WinJS.Binding.converter(function (onSale) { return onSale ? "block" : "none"; }) }); app.start(); })(); The ListViewDemos.displayNormalProduct binding converter converts the value true or false to the value “none” or “block”. The ListViewDemos.displayOnSaleProduct binding converter does the opposite; it converts the value true or false to the value “block” or “none” (Sadly, you cannot simply place a NOT operator before the onSale property in the binding expression – you need to create both converters). The end result is that you can display different markup depending on the value of the product onSale property. Either the contents of the first or second DIV element are displayed: Summary In this blog entry, I’ve explored two approaches to displaying different markup in a ListView depending on the value of a data item property. The bulk of this blog entry was devoted to explaining how you can assign a function to the ListView itemTemplate property which returns different templates. We created both a productItemTemplate and productOnSaleTemplate and displayed both templates with the same ListView control. We also discussed how you can create a single template and display different markup by using binding converters. The binding converters are used to set a DIV element’s display property to either “none” or “block”. We created a binding converter which displays normal products and a binding converter which displays “on sale” products.

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  • Web API, JavaScript, Chrome &amp; Cross-Origin Resource Sharing

    - by Brian Lanham
    The team spent much of the week working through this issues related to Chrome running on Windows 8 consuming cross-origin resources using Web API.  We thought it was resolved on day 2 but it resurfaced the next day.  We definitely resolved it today though.  I believe I do not fully understand the situation but I am going to explain what I know in an effort to help you avoid and/or resolve a similar issue. References We referenced many sources during our trial-and-error troubleshooting.  These are the links we reference in order of applicability to the solution: Zoiner Tejada JavaScript and other material from -> http://www.devproconnections.com/content1/topic/microsoft-azure-cors-141869/catpath/windows-azure-platform2/page/3 WebDAV Where I learned about “Accept” –>  http://www-jo.se/f.pfleger/cors-and-iis? IT Hit Tells about NOT using ‘*’ –> http://www.webdavsystem.com/ajax/programming/cross_origin_requests Carlos Figueira Sample back-end code (newer) –> http://code.msdn.microsoft.com/windowsdesktop/Implementing-CORS-support-a677ab5d (older version) –> http://code.msdn.microsoft.com/CORS-support-in-ASPNET-Web-01e9980a   Background As a measure of protection, Web designers (W3C) and implementers (Google, Microsoft, Mozilla) made it so that a request, especially a JSON request (but really any URL), sent from one domain to another will only work if the requestee “knows” about the requester and allows requests from it. So, for example, if you write a ASP.NET MVC Web API service and try to consume it from multiple apps, the browsers used may (will?) indicate that you are not allowed by showing an “Access-Control-Allow-Origin” error indicating the requester is not allowed to make requests. Internet Explorer (big surprise) is the odd-hair-colored step-child in this mix. It seems that running locally at least IE allows this for development purposes.  Chrome and Firefox do not.  In fact, Chrome is quite restrictive.  Notice the images below. IE shows data (a tabular view with one row for each day of a week) while Chrome does not (trust me, neither does Firefox).  Further, the Chrome developer console shows an XmlHttpRequest (XHR) error. Screen captures from IE (left) and Chrome (right). Note that Chrome does not display data and the console shows an XHR error. Why does this happen? The Web browser submits these requests and processes the responses and each browser is different. Okay, so, IE is probably the only one that’s truly different.  However, Chrome has a specific process of performing a “pre-flight” check to make sure the service can respond to an “Access-Control-Allow-Origin” or Cross-Origin Resource Sharing (CORS) request.  So basically, the sequence is, if I understand correctly:  1)Page Loads –> 2)JavaScript Request Processed by Browser –> 3)Browsers Prepares to Submit Request –> 4)[Chrome] Browser Submits Pre-Flight Request –> 5)Server Responds with HTTP 200 –> 6)Browser Submits Request –> 7)Server Responds with Data –> 8)Page Shows Data This situation occurs for both GET and POST methods.  Typically, GET methods are called with query string parameters so there is no data posted.  Instead, the requesting domain needs to be permitted to request data but generally nothing more is required.  POSTs on the other hand send form data.  Therefore, more configuration is required (you’ll see the configuration below).  AJAX requests are not friendly with this (POSTs) either because they don’t post in a form. How to fix it. The team went through many iterations of self-hair removal and we think we finally have a working solution.  The trial-and-error approach eventually worked and we referenced many sources for the information.  I indicate those references above.  There are basically three (3) tasks needed to make this work. Assumptions: You are using Visual Studio, Web API, JavaScript, and have Cross-Origin Resource Sharing, and several browsers. 1. Configure the client Joel Cochran centralized our “cors-oriented” JavaScript (from here). There are two calls including one for GET and one for POST function(url, data, callback) {             console.log(data);             $.support.cors = true;             var jqxhr = $.post(url, data, callback, "json")                 .error(function(jqXhHR, status, errorThrown) {                     if ($.browser.msie && window.XDomainRequest) {                         var xdr = new XDomainRequest();                         xdr.open("post", url);                         xdr.onload = function () {                             if (callback) {                                 callback(JSON.parse(this.responseText), 'success');                             }                         };                         xdr.send(data);                     } else {                         console.log(">" + jqXhHR.status);                         alert("corsAjax.post error: " + status + ", " + errorThrown);                     }                 });         }; The GET CORS JavaScript function (credit to Zoiner Tejada) function(url, callback) {             $.support.cors = true;             var jqxhr = $.get(url, null, callback, "json")                 .error(function(jqXhHR, status, errorThrown) {                     if ($.browser.msie && window.XDomainRequest) {                         var xdr = new XDomainRequest();                         xdr.open("get", url);                         xdr.onload = function () {                             if (callback) {                                 callback(JSON.parse(this.responseText), 'success');                             }                         };                         xdr.send();                     } else {                         alert("CORS is not supported in this browser or from this origin.");                     }                 });         }; The POST CORS JavaScript function (credit to Zoiner Tejada) Now you need to call these functions to get and post your data (instead of, say, using $.Ajax). Here is a GET example: corsAjax.get(url, function(data) { if (data !== null && data.length !== undefined) { // do something with data } }); And here is a POST example: corsAjax.post(url, item); Simple…except…you’re not done yet. 2. Change Web API Controllers to Allow CORS There are actually two steps here.  Do you remember above when we mentioned the “pre-flight” check?  Chrome actually asks the server if it is allowed to ask it for cross-origin resource sharing access.  So you need to let the server know it’s okay.  This is a two-part activity.  a) Add the appropriate response header Access-Control-Allow-Origin, and b) permit the API functions to respond to various methods including GET, POST, and OPTIONS.  OPTIONS is the method that Chrome and other browsers use to ask the server if it can ask about permissions.  Here is an example of a Web API controller thus decorated: NOTE: You’ll see a lot of references to using “*” in the header value.  For security reasons, Chrome does NOT recognize this is valid. [HttpHeader("Access-Control-Allow-Origin", "http://localhost:51234")] [HttpHeader("Access-Control-Allow-Credentials", "true")] [HttpHeader("Access-Control-Allow-Methods", "ACCEPT, PROPFIND, PROPPATCH, COPY, MOVE, DELETE, MKCOL, LOCK, UNLOCK, PUT, GETLIB, VERSION-CONTROL, CHECKIN, CHECKOUT, UNCHECKOUT, REPORT, UPDATE, CANCELUPLOAD, HEAD, OPTIONS, GET, POST")] [HttpHeader("Access-Control-Allow-Headers", "Accept, Overwrite, Destination, Content-Type, Depth, User-Agent, X-File-Size, X-Requested-With, If-Modified-Since, X-File-Name, Cache-Control")] [HttpHeader("Access-Control-Max-Age", "3600")] public abstract class BaseApiController : ApiController {     [HttpGet]     [HttpOptions]     public IEnumerable<foo> GetFooItems(int id)     {         return foo.AsEnumerable();     }     [HttpPost]     [HttpOptions]     public void UpdateFooItem(FooItem fooItem)     {         // NOTE: The fooItem object may or may not         // (probably NOT) be set with actual data.         // If not, you need to extract the data from         // the posted form manually.         if (fooItem.Id == 0) // However you check for default...         {             // We use NewtonSoft.Json.             string jsonString = context.Request.Form.GetValues(0)[0].ToString();             Newtonsoft.Json.JsonSerializer js = new Newtonsoft.Json.JsonSerializer();             fooItem = js.Deserialize<FooItem>(new Newtonsoft.Json.JsonTextReader(new System.IO.StringReader(jsonString)));         }         // Update the set fooItem object.     } } Please note a few specific additions here: * The header attributes at the class level are required.  Note all of those methods and headers need to be specified but we find it works this way so we aren’t touching it. * Web API will actually deserialize the posted data into the object parameter of the called method on occasion but so far we don’t know why it does and doesn’t. * [HttpOptions] is, again, required for the pre-flight check. * The “Access-Control-Allow-Origin” response header should NOT NOT NOT contain an ‘*’. 3. Headers and Methods and Such We had most of this code in place but found that Chrome and Firefox still did not render the data.  Interestingly enough, Fiddler showed that the GET calls succeeded and the JSON data is returned properly.  We learned that among the headers set at the class level, we needed to add “ACCEPT”.  Note that I accidentally added it to methods and to headers.  Adding it to methods worked but I don’t know why.  We added it to headers also for good measure. [HttpHeader("Access-Control-Allow-Methods", "ACCEPT, PROPFIND, PROPPA... [HttpHeader("Access-Control-Allow-Headers", "Accept, Overwrite, Destin... Next Steps That should do it.  If it doesn’t let us know.  What to do next?  * Don’t hardcode the allowed domains.  Note that port numbers and other domain name specifics will cause problems and must be specified.  If this changes do you really want to deploy updated software?  Consider Miguel Figueira’s approach in the following link to writing a custom HttpHeaderAttribute class that allows you to specify the domain names and then you can do it dynamically.  There are, of course, other ways to do it dynamically but this is a clean approach. http://code.msdn.microsoft.com/windowsdesktop/Implementing-CORS-support-a677ab5d

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