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  • Ajax Control Toolkit November 2011 Release

    - by Stephen Walther
    I’m happy to announce the November 2011 Release of the Ajax Control Toolkit. This release introduces a new Balloon Popup control and several enhancements to the existing Tabs control including support for on-demand loading of tab content, support for vertical tabs, and support for keyboard tab navigation. We also fixed the top-voted bugs associated with the Tabs control reported at CodePlex.com. You can download the new release by visiting the CodePlex website: http://AjaxControlToolkit.CodePlex.com Alternatively, the fast and easy way to get the latest release of the Ajax Control Toolkit is to use NuGet. Open your Library Package Manager console in Visual Studio 2010 and type: After you install the Ajax Control Toolkit through NuGet, please do a Rebuild of your project (the menu option Build, Rebuild). After you do a Rebuild, the ajaxToolkit prefix will appear in Intellisense: Using the Balloon Popup Control Why a new Balloon Popup control? The Balloon Popup control is the second most requested new feature for the Ajax Control Toolkit according to CodePlex votes: The Balloon Popup displays a message in a balloon when you shift focus to a control, click a control, or hover over a control. You can use the Balloon Popup, for example, to display instructions for TextBoxes which appear in a form: Here’s the code used to create the Balloon Popup: <ajaxToolkit:ToolkitScriptManager ID="tsm1" runat="server" /> <asp:TextBox ID="txtFirstName" Runat="server" /> <asp:Panel ID="pnlFirstNameHelp" runat="server"> Please enter your first name </asp:Panel> <ajaxToolkit:BalloonPopupExtender TargetControlID="txtFirstName" BalloonPopupControlID="pnlFirstNameHelp" BalloonSize="Small" UseShadow="true" runat="server" /> You also can use the Balloon Popup to explain hard to understand words in a text document: Here’s how you display the Balloon Popup when you hover over the link: The point of the conversation was <asp:HyperLink ID="lnkObfuscate" Text="obfuscated" CssClass="hardWord" runat="server" /> by his incessant coughing. <ajaxToolkit:ToolkitScriptManager ID="tsm1" runat="server" /> <asp:Panel id="pnlObfuscate" Runat="server"> To bewilder or render something obscure </asp:Panel> <ajaxToolkit:BalloonPopupExtender TargetControlID="lnkObfuscate" BalloonPopupControlID="pnlObfuscate" BalloonStyle="Cloud" UseShadow="true" DisplayOnMouseOver="true" Runat="server" />   There are four important properties which you need to know about when using the Balloon Popup control: BalloonSize – The three balloon sizes are Small, Medium, and Large. BalloonStyle -- The two built-in styles are Rectangle and Cloud. UseShadow – When true, a drop shadow appears behind the popup. Position – Can have the values Auto, BottomLeft, BottomRight, TopLeft, TopRight. When set to Auto, which is the default, the Balloon Popup will appear where it has the most screen real estate. The following screenshots illustrates how these settings affect the appearance of the Balloon Popup: Customizing the Balloon Popup You can customize the appearance of the Balloon Popup by creating your own Cascading Style Sheet and Sprite. The Ajax Control Toolkit sample site includes a sample of a custom Oval Balloon Popup style: This custom style was created by using a custom Cascading Style Sheet and image. You point the Balloon Popup at a custom Cascading Style Sheet and Cascading Style Sheet class by using the CustomCssUrl and CustomClassName properties like this: <asp:TextBox ID="txtCustom" autocomplete="off" runat="server" /> <br /> <asp:Panel ID="Panel3" runat="server"> This is a custom BalloonPopupExtender style created with a custom Cascading Style Sheet. </asp:Panel> <ajaxToolkit:BalloonPopupExtender ID="bpe1" TargetControlID="txtCustom" BalloonPopupControlID="Panel3" BalloonStyle="Custom" CustomCssUrl="CustomStyle/BalloonPopupOvalStyle.css" CustomClassName="oval" UseShadow="true" runat="server" />   Learn More about the Balloon Popup To learn more about the Balloon Popup control, visit the sample page for the Balloon Popup at the Ajax Control Toolkit sample site: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/BalloonPopup/BalloonPopupExtender.aspx Improvements to the Tabs Control In this release, we introduced several important new features for the existing Tabs control. We also fixed all of the top-voted bugs for the Tabs control. On-Demand Loading of Tab Content Here is the scenario. Imagine that you are using the Tabs control in a Web Forms page. The Tabs control displays two tabs: Customers and Products. When you click the Customers tab then you want to see a list of customers and when you click on the Products tab then you want to see a list of products. In this scenario, you don’t want the list of customers and products to be retrieved from the database when the page is initially opened. The user might never click on the Products tab and all of the work to load the list of products from the database would be wasted. In this scenario, you want the content of a tab panel to be loaded on demand. The products should only be loaded from the database and rendered to the browser when you click the Products tab and not before. The Tabs control in the November 2011 Release of the Ajax Control Toolkit includes a new property named OnDemand. When OnDemand is set to the value True, a tab panel won’t be loaded until you click its associated tab. Here is the code for the aspx page: <ajaxToolkit:ToolkitScriptManager ID="tsm1" runat="server" /> <ajaxToolkit:TabContainer ID="tabs" OnDemand="false" runat="server"> <ajaxToolkit:TabPanel HeaderText="Customers" runat="server"> <ContentTemplate> <h2>Customers</h2> <asp:GridView ID="grdCustomers" DataSourceID="srcCustomers" runat="server" /> <asp:SqlDataSource ID="srcCustomers" SelectCommand="SELECT * FROM Customers" ConnectionString="<%$ ConnectionStrings:StoreDB %>" runat="server" /> </ContentTemplate> </ajaxToolkit:TabPanel> <ajaxToolkit:TabPanel HeaderText="Products" runat="server"> <ContentTemplate> <h2>Products</h2> <asp:GridView ID="grdProducts" DataSourceID="srcProducts" runat="server" /> <asp:SqlDataSource ID="srcProducts" SelectCommand="SELECT * FROM Products" ConnectionString="<%$ ConnectionStrings:StoreDB %>" runat="server" /> </ContentTemplate> </ajaxToolkit:TabPanel> </ajaxToolkit:TabContainer> Notice that the TabContainer includes an OnDemand=”True” property. The Tabs control contains two Tab Panels. The first tab panel uses a DataGrid and SqlDataSource to display a list of customers and the second tab panel uses a DataGrid and SqlDataSource to display a list of products. And here is the code-behind for the page: using System; using System.Diagnostics; using System.Web.UI.WebControls; namespace ACTSamples { public partial class TabsOnDemand : System.Web.UI.Page { protected override void OnInit(EventArgs e) { srcProducts.Selecting += new SqlDataSourceSelectingEventHandler(srcProducts_Selecting); } void srcProducts_Selecting(object sender, SqlDataSourceSelectingEventArgs e) { Debugger.Break(); } } } The code-behind file includes an event handler for the Products SqlDataSource Selecting event. The handler breaks into the debugger by calling the Debugger.Break() method. That way, we can know when the Products SqlDataSource actually retrieves the list of products. When the OnDemand property has the value False then the Selecting event handler is called immediately when the page is first loaded. The contents of all of the tabs are loaded (and the contents of the unselected tabs are hidden) when the page is first loaded. When the OnDemand property has the value True then the Selecting event handler is not called when the page is first loaded. The event handler is not called until you click on the Products tab. If you never click on the Products tab then the list of products is never retrieved from the database. If you want even more control over when the contents of a tab panel gets loaded then you can use the TabPanel OnDemandMode property. This property accepts the following three values: None – Never load the contents of the tab panel again after the page is first loaded. Once – Wait until the tab is selected to load the contents of the tab panel Always – Load the contents of the tab panel each and every time you select the tab. There is a live demonstration of the OnDemandMode property here in the sample page for the Tabs control: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/Tabs/Tabs.aspx Displaying Vertical Tabs With the November 2011 Release, the Tabs control now supports vertical tabs. To create vertical tabs, just set the TabContainer UserVerticalStripPlacement property to the value True like this: <ajaxToolkit:TabContainer ID="tabs" OnDemand="false" UseVerticalStripPlacement="true" runat="server"> <ajaxToolkit:TabPanel ID="TabPanel1" HeaderText="First Tab" runat="server"> <ContentTemplate> <p> Lorem ipsum dolor sit amet, consectetuer adipiscing elit. Maecenas porttitor congue massa. Fusce posuere, magna sed pulvinar ultricies, purus lectus malesuada libero, sit amet commodo magna eros quis urna. </p> </ContentTemplate> </ajaxToolkit:TabPanel> <ajaxToolkit:TabPanel ID="TabPanel2" HeaderText="Second Tab" runat="server"> <ContentTemplate> <p> Lorem ipsum dolor sit amet, consectetuer adipiscing elit. Maecenas porttitor congue massa. Fusce posuere, magna sed pulvinar ultricies, purus lectus malesuada libero, sit amet commodo magna eros quis urna. </p> </ContentTemplate> </ajaxToolkit:TabPanel> </ajaxToolkit:TabContainer> In addition, you can use the TabStripPlacement property to control whether the tab strip appears at the left or right or top or bottom of the tab panels: Tab Keyboard Navigation Another highly requested feature for the Tabs control is support for keyboard navigation. The Tabs control now supports the arrow keys and the Home and End keys. In order for the arrow keys to work, you must first move focus to the tab control on the page by either clicking on a tab with your mouse or repeatedly hitting the Tab key. You can try out the new keyboard navigation support by trying any of the demos included in the Tabs sample page: http://www.asp.net/ajaxLibrary/AjaxControlToolkitSampleSite/Tabs/Tabs.aspx Summary I hope that you take advantage of the new Balloon Popup control and the new features which we introduced for the Tabs control. We added a lot of new features to the Tabs control in this release including support for on-demand tabs, support for vertical tabs, and support for tab keyboard navigation. I want to thank the developers on the Superexpert team for all of the hard work which they put into this release.

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  • September 2012 Release of the Ajax Control Toolkit

    - by Stephen.Walther
    I’m excited to announce the September 2012 release of the Ajax Control Toolkit! This is the first release of the Ajax Control Toolkit which supports the .NET 4.5 framework. We also continue to support ASP.NET 3.5 and ASP.NET 4.0. With this release, we’ve made several important bug fixes. The Superexpert team focused on fixing the highest voted issues associated with the CascadingDropDown control. I’ve created a list of these bug fixes later in this blog post. You can download the latest release of the Ajax Control Toolkit by visiting the following page at CodePlex: http://AjaxControlToolkit.CodePlex.com Alternatively, you can install the latest version of the Ajax Control Toolkit using NuGet by firing off the following command from the Package Manager Console: Install-Package AjaxControlToolkit Using the Ajax Control Toolkit with ASP.NET 4.5 Let me walk through the steps for using the Ajax Control Toolkit with ASP.NET 4.5. First, I’ll create a new ASP.NET 4.5 website with Visual Studio 2012. I’ll create the new website with the ASP.NET Web Forms Application template: When you create a new ASP.NET 4.5 site with the ASP.NET Web Forms Application template, you get a starter website. If you run the site, then you get a page with default content: Let me show you how you can add the Ajax Control Toolkit Calendar control to the homepage of this starter site. The first step is to use NuGet to install the Ajax Control Toolkit. Right-click the References folder in the Solution Explorer window and select the menu option Manage NuGet Packages. In the Manage NuGet Packages dialog, use the search box to search for the Ajax Control Toolkit (enter “AjaxControlToolkit”). After you find it, click the Install button to add the Ajax Control Toolkit to your project. That’s all you have to do to install the Ajax Control Toolkit! Now we are ready to start using the Ajax Control Toolkit controls. Open the default.aspx page so we can modify the contents of the page. Erase everything contained in the Content control with the ID of BodyContent. After erasing the content, declare the following two controls: <asp:TextBox ID="vacationDate" runat="server" /> <ajaxToolkit:CalendarExtender TargetControlID="vacationDate" runat="server" /> The first control is a standard ASP.NET TextBox control and the second control is an Ajax Control Toolkit Calendar control. You should get intellisense as you type out the Ajax Control Toolkit Calendar control. If you don’t, then close and re-open the Default.aspx page. Now, let’s run our app. Hit the F5 button or select Debug, Start Debugging from the Visual Studio menu. You will get the error message “MsAjaxBundle is not a valid script name”. Don’t despair! We need to update the Master Page so it uses the ToolkitScriptManager instead of the default ScriptManager. Open the Site.Master file and find where the ScriptManager is declared. The ScriptManager should look like this: <asp:ScriptManager runat="server"> <Scripts> <%--Framework Scripts--%> <asp:ScriptReference Name="MsAjaxBundle" /> <asp:ScriptReference Name="jquery" /> <asp:ScriptReference Name="jquery.ui.combined" /> <asp:ScriptReference Name="WebForms.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebForms.js" /> <asp:ScriptReference Name="WebUIValidation.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebUIValidation.js" /> <asp:ScriptReference Name="MenuStandards.js" Assembly="System.Web" Path="~/Scripts/WebForms/MenuStandards.js" /> <asp:ScriptReference Name="GridView.js" Assembly="System.Web" Path="~/Scripts/WebForms/GridView.js" /> <asp:ScriptReference Name="DetailsView.js" Assembly="System.Web" Path="~/Scripts/WebForms/DetailsView.js" /> <asp:ScriptReference Name="TreeView.js" Assembly="System.Web" Path="~/Scripts/WebForms/TreeView.js" /> <asp:ScriptReference Name="WebParts.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebParts.js" /> <asp:ScriptReference Name="Focus.js" Assembly="System.Web" Path="~/Scripts/WebForms/Focus.js" /> <asp:ScriptReference Name="WebFormsBundle" /> <%--Site Scripts--%> </Scripts> </asp:ScriptManager> We need to make three changes to the ScriptManager: 1) We need to replace the asp:ScriptManager with the ajaxToolkit:ToolkitScriptManager 2) We need to remove the MsAjaxBundle bundle from the ScriptReferences 3) We need to remove the Assembly=”System.Web” attributes from the ScriptReferences After you make these three changes, the ToolkitScriptManager should looks like this: <ajaxToolkit:ToolkitScriptManager runat="server"> <Scripts> <%--Framework Scripts--%> <asp:ScriptReference Name="jquery" /> <asp:ScriptReference Name="jquery.ui.combined" /> <asp:ScriptReference Name="WebForms.js" Path="~/Scripts/WebForms/WebForms.js" /> <asp:ScriptReference Name="WebUIValidation.js" Path="~/Scripts/WebForms/WebUIValidation.js" /> <asp:ScriptReference Name="MenuStandards.js" Path="~/Scripts/WebForms/MenuStandards.js" /> <asp:ScriptReference Name="GridView.js" Path="~/Scripts/WebForms/GridView.js" /> <asp:ScriptReference Name="DetailsView.js" Path="~/Scripts/WebForms/DetailsView.js" /> <asp:ScriptReference Name="TreeView.js" Path="~/Scripts/WebForms/TreeView.js" /> <asp:ScriptReference Name="WebParts.js" Path="~/Scripts/WebForms/WebParts.js" /> <asp:ScriptReference Name="Focus.js" Path="~/Scripts/WebForms/Focus.js" /> <asp:ScriptReference Name="WebFormsBundle" /> <%--Site Scripts--%> </Scripts> </ajaxToolkit:ToolkitScriptManager> After we make these changes, the app should run successfully. You’ll get a page which contains a text field. When you click inside the text field, a popup calendar is displayed. Ajax Control Toolkit and jQuery You might have noticed that the ScriptManager includes a reference to jQuery by default. We did not remove that reference when we converted the ScriptManager to a ToolkitScriptManager. You can use the Ajax Control Toolkit and jQuery side-by-side. Here’s how you can modify the Default.aspx page so that it contains two popup calendars. The first popup calendar is created with the Ajax Control Toolkit and the second popup calendar is created with jQuery: <asp:TextBox ID="vacationDate" runat="server" /> <ajaxToolkit:CalendarExtender TargetControlID="vacationDate" runat="server" /> <input id="birthDate" /> <script> $("#birthDate").datepicker(); </script> Before you can start using jQuery UI plugins, you need to complete one more step. You need to add the jQuery UI themes bundle to the HEAD of the Site.Master page like this: <head runat="server"> <meta charset="utf-8" /> <title><%: Page.Title %> - My ASP.NET Application</title> <asp:PlaceHolder runat="server"> <%: Scripts.Render("~/bundles/modernizr") %> </asp:PlaceHolder> <webopt:BundleReference runat="server" Path="~/Content/css" /> <webopt:BundleReference runat="server" Path="~/Content/themes/base/css" /> <link href="~/favicon.ico" rel="shortcut icon" type="image/x-icon" /> <meta name="viewport" content="width=device-width" /> <asp:ContentPlaceHolder runat="server" ID="HeadContent" /> </head> The markup above includes a reference to the jQuery UI themes bundle: <webopt:BundleReference runat="server" Path="~/Content/themes/base/css" /> Now that we have made these changes, we can use the Ajax Control Toolkit and jQuery at the same time. When you run your app, you get two popup calendars. When you click in the first text field, the Ajax Control Toolkit calendar appears. When you click in the second text field, the jQuery UI popup calendar appears: Bug Fixes in this Release We made several important bug fixes with this release of the Ajax Control Toolkit and integrated several Pull Requests contributed by the community. Our primary focus during this sprint was fixing issues with the CascadingDropDown control. We fixed the following issues associated with the CascadingDropDown: · 9490 – Don’t disable dropdowns in CascadingDropDown · 14223 – CascadingDropDown Reset or Setting SelectedValue from WebMethod · 12189 – CascadingDropDown not obeying disabled state of DropDownList · 22942 – CascadingDropDown infinite loop (with solution) · 8671 – CascadingDropdown options is null or undefined · 14407 – CascadingDropDown: populated client event happens too often · 17148 – CascadingDropDown – Add “UseHttpGet” property · 10221 – No NotNull check in CascadingDropDown · 12228 – Provide property for case-insensitive DefaultValue lookup in CascadingDropdown We also fixed the following two issues which are not directly related to the CascadingDropDown control: · 27108 – CalendarExtender: Bug when selecting December shifts to January. · 27041 – Input controls with HTML5 types do not post back in Firefox, Chrome, Safari Finally, we integrated several Pull Requests submitted by the community (Thank you community!): · Added French localized resources for the AjaxFileUpload · Resolved an issue which prevented the AjaxFileUpload control from working with pages that require query string variables. · Extended the AjaxFileUploadEventArgs class to include the current file index in the queue and the total number of files in the queue. · Fixed an issue with TabContainer and TabPanel which caused the OnActiveTabChanged event to fire too often. Summary I’m happy to see the Ajax Control Toolkit move forward into the brave new world of ASP.NET 4.5! In this latest release, we focused on ensuring that the Ajax Control Toolkit works smoothly with ASP.NET 4.5 applications. We also fixed the highest voted bugs associated with the CascadingDropDown control and integrated several Pull Request submitted by the community. Once again, I want to thank the Superexpert team for their hard work on this release!

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  • Why does this mod_rewrite rule 'not-match'? (big rewrite log included)

    - by Christopher
    I've got a scenario involving two domains: WordPress site hosted on domain1.com domain2.co.uk, simply redirecting users to domain1 via mod_rewrite This rule applies irrespective of whether www. is specified or not. (It's eventually removed from the URL, I'm a no-WWW fan.) There's nothing on domain2.co.uk at all except for an .htaccess with some mod_rewrite rules. However, I want to be able to allow users to be redirected to the correct article URI even if they specify the "wrong" URL (i.e., a 301 redirect preserving the stuff after the first forward slash). I'm currently achieving this with this ruleset: RewriteCond %{HTTP_HOST} ^((www\.)?[^\.]+)\.domain2\.co\.uk [NC,OR] RewriteCond %{HTTP_HOST} ^domain2\.co\.uk [NC] RewriteRule ^(.*)$ http://domain1.com/$1 [R=301,L] This works but is uglier than I want it to be. I'm not a mod_rewrite zen master, but from what I can tell the top rule should match irrespective of whether www. is specified... But it doesn't. In order to catch www-less requests, I need the second RewriteCond. From the rewrite log, with just the first RewriteCond: [domain2.co.uk/sid#e200498][rid#e670168/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] strip per-dir prefix: /home/devnull/domains/domain2.co.uk/public_html/ -> [domain2.co.uk/sid#e200498][rid#e670168/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] applying pattern '^(.*)$' to uri '' [domain2.co.uk/sid#e200498][rid#e670168/initial] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^((www\.)|[^\.]+)\.domain2\.co\.uk' [NC] => not-matched [domain2.co.uk/sid#e200498][rid#e670168/initial] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/ [domain2.co.uk/sid#e200498][rid#e653868/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.html [domain2.co.uk/sid#e200498][rid#e65f8b8/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.htm [domain2.co.uk/sid#e200498][rid#e653868/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.shtml [domain2.co.uk/sid#e200498][rid#e65f8b8/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.php [domain2.co.uk/sid#e200498][rid#e653868/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.php5 [domain2.co.uk/sid#e200498][rid#e666c98/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.php4 [domain2.co.uk/sid#e200498][rid#e65f8b8/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.php3 [domain2.co.uk/sid#e200498][rid#e653868/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.phtml [domain2.co.uk/sid#e200498][rid#e65f8b8/subreq] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/index.cgi [domain2.co.uk/sid#e200498][rid#e66c370/initial/redir#1] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] strip per-dir prefix: /home/devnull/domains/domain2.co.uk/public_html/403.shtml -> 403.shtml [domain2.co.uk/sid#e200498][rid#e66c370/initial/redir#1] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] applying pattern '^(.*)$' to uri '403.shtml' [domain2.co.uk/sid#e200498][rid#e66c370/initial/redir#1] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^((www\.)|[^\.]+)\.domain2\.co\.uk' [NC] => not-matched [domain2.co.uk/sid#e200498][rid#e66c370/initial/redir#1] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/403.shtml [domain2.co.uk/sid#e200498][rid#e668ca8/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] strip per-dir prefix: /home/devnull/domains/domain2.co.uk/public_html/favicon.ico -> favicon.ico [domain2.co.uk/sid#e200498][rid#e668ca8/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] applying pattern '^(.*)$' to uri 'favicon.ico' [domain2.co.uk/sid#e200498][rid#e668ca8/initial] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^((www\.)|[^\.]+)\.domain2\.co\.uk' [NC] => not-matched [domain2.co.uk/sid#e200498][rid#e668ca8/initial] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/favicon.ico [domain2.co.uk/sid#e200498][rid#f160b40/initial/redir#1] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] strip per-dir prefix: /home/devnull/domains/domain2.co.uk/public_html/404.shtml -> 404.shtml [domain2.co.uk/sid#e200498][rid#f160b40/initial/redir#1] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] applying pattern '^(.*)$' to uri '404.shtml' [domain2.co.uk/sid#e200498][rid#f160b40/initial/redir#1] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^((www\.)|[^\.]+)\.domain2\.co\.uk' [NC] => not-matched [domain2.co.uk/sid#e200498][rid#f160b40/initial/redir#1] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] pass through /home/devnull/domains/domain2.co.uk/public_html/404.shtml However with the second RewriteCond added, the rule works, and the logs show this: [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] strip per-dir prefix: /home/devnull/domains/domain2.co.uk/public_html/ -> [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (3) [perdir /home/devnull/domains/domain2.co.uk/public_html/] applying pattern '^(.*)$' to uri '' [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^((www\.)?[^\.]+)\.domain2\.co\.uk' [NC] => not-matched [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (4) [perdir /home/devnull/domains/domain2.co.uk/public_html/] RewriteCond: input='domain2.co.uk' pattern='^domain2\.co\.uk' [NC] => matched [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (2) [perdir /home/devnull/domains/domain2.co.uk/public_html/] rewrite '' -> 'http://domain1.com/' [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (2) [perdir /home/devnull/domains/domain2.co.uk/public_html/] explicitly forcing redirect with http://domain1.com/ [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] escaping http://domain1.com/ for redirect [domain2.co.uk/sid#e200498][rid#e65fe58/initial] (1) [perdir /home/devnull/domains/domain2.co.uk/public_html/] redirect to http://domain1.com/ [REDIRECT/301] Can anybody help me figure out why it just won't work with the one rule? I feel like I'm missing the bleeding obvious, and while the second RewriteCond is a valid workaround, it's a kludge and that annoys me. ;-) All help appreciated...

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  • Begin the Clone Wars Have!

    - by Antony Reynolds
    Creating a New Virtual Machine from an Existing Virtual Disk In previous posts I described how I set up an OEL6 machine under VirtualBox that can run an 11gR2 database and FMW 11.1.1.5.  That is great if you want the DB and FMW running in the same virtual image and it has served me well for some proof of concepts and also for some testing of different JVMs.  However I also wanted to run some testing of FMW with the database running on a separate physical machine.  So in this post I will show how to take a VirtualBox image and create a new image based on the disks from that original image. What are my Options? There is more than one way to skin a cat, or in this case to create two separate VMs that can run on different hardware.  Some of the options include: Create new virtual disk images for each new VM. Clone the existing disk images and point the new VM at the cloned images. Point the new VM at the existing snapshots. #1 is too much like hard work, install OEL twice, install a database again, install FMW again, run RCU again!  Life is too short! #2 is probably the safest way of doing things.  VirtualBox allows you to clone a disk image for use in a separate machine.  However this of course duplicates the disk and means that it is now occupying 3 times the space, once for the original disk and twice more for the two clones I would need. #3 is the most space efficient way of doing things.  It does mean however that I can only run the new “cloned” images if I have access to the original image because that is where the base snapshots reside.  However this is not a problem for me as long as I remember to keep all threee images together.  So this is the approach we will follow. Snapshot, What Snapshot? As we are going to create new virtual machines based on existing snapshots we need to figure out which snapshot to use.  We do this by opening the “Media Manager” from within VirtualBox and moving the mouse over the snapshot images until we find the snapshots we want – the snapshot name is identified in the “Attached to:” comment.  In my case I wanted the FMW installed snapshot because that had a database configured for FMW alongside the FMW software.  I made a note of the filename of that snapshot (actually I just noted the first 5 characters as that was all that was needed to uniquely identify the snapshot file). When we create the new machines we will point them at the snapshot filename we have just checked. Network or NotWork? Because we want the two new machines to communicate with each other when hosted in different physical machines we can’t use the default NAT networking mode without a lot of hassle.  But at the same time we need them to have fixed IP addresses relative to each other so that they can see each other whilst also being able to see the outside world. To achieve all these requirements I created two network adapters for each machine.  Adapter 1 was a standard NAT mapping.  This will allow each machine to get a dynamic IP address (10.0.2.15 by default) that can be used to access the external world through the VBox provided NAT gateway.  This is the same as the existing configuration. The second adapter I created as a bridged adapter.  This gives the virtual machine direct access to the host network card and by using fixed IP addresses each machine can see the other.  It is important to choose fixed IP addresses that are not routable across your internal network so you don’t get any clashes with other machines on your network.  Of course you could always get proper fixed IP addresses from your network people, but I have serveral people using my images and as long as I don’t have two instances of the same VM on the same network segment this is easier and avoids reconfiguring the network every time someone wants a copy of my VM.  If it is available I would suggest using the 10.0.3.* network as 10.0.2.* is the default NAT network.  You can check availability by pinging 10.0.3.1 and 10.0.3.2 from your host machine.  If it times out then you are probably safe to use that. Creating the New VMs Now that I had collected the data that I needed I went ahead and created the new VMs. When asked for a “Boot Hard Disk” I used the “Choose a virtual hard disk file…” link to find the snapshot I had previously selected and set that to be the existing hard disk.  I chose the previously existing SOA 11.1.1.5 install for both the new DB and FMW machines because that snapshot had the database with the RCU completed that I wanted for my DB machine and it had the SOA software installed which I wanted for my FMW machine. After the initial creation of the virtual machine go into the network setting section and enable a second adapter which will be bridged.  Make a note of the MAC addresses (the last four digits should be sufficient) of the two adapters so that you can later set the bridged adapter to use fixed IP and the NAT adapter to use DHCP. We are now ready to start the VMs and reconfigure Linux. Reconfiguring Linux Because I now have two new machines I need to change their network configuration.  In particular I need to change the hostname, update the hosts file and change the network settings. Changing the Hostname I renamed both hosts by running the hostname command as root: hostname vboxfmw.oracle.com I also edited the /etc/sysconfig file and set the correct hostname in there. HOSTNAME=vboxfmw.oracle.com Changing the Network Settings I needed to change the network configuration to give the bridged network a fixed IP address.  I first explicitly set the MAC addresses of the two adapters, because the order of the virtual adapters in the VirtualBox Manager is not necessarily the same as the order of the adapters in the guest OS.  So I went in to the System->Preferences->Network Connections screen and explicitly set the “Device MAC address” for the two adapters. Having correctly mapped the Linux adapters to the VirtualBox adapters I then set the Bridged adapter to use fixed IP addressing rather than DHCP.  There is no need for additional routing or default gateways because we expect the two machine to be on the same LAN segment. Updating the Hosts File Having renamed the machines and reconfigured the network I then updated the /etc/hosts file to refer to the new machine name add a new line to the hosts file to provide an additional IP address for my server (the new fixed IP address) add a new line for the fixed IP address of the other virtual machine 10.0.3.101      vboxdb.oracle.com       vboxdb  # Added by NetworkManager 10.0.2.15       vboxdb.oracle.com       vboxdb  # Added by NetworkManager 10.0.3.102      vboxfmw.oracle.com      vboxfmw # Added by NetworkManager 127.0.0.1       localhost.localdomain   localhost ::1     vboxdb.oracle.com       vboxdb  localhost6.localdomain6 localhost6 To make sure everything takes effect I restarted the server. Reconfiguring the Database on the DB Machine Because we changed the hostname the listener and the EM console no longer start so I need to modify the listener.ora to use the new hostname and I also need to rebuild the EM configuration because it also relies on the hostname. I edited the $ORACLE_HOME/network/admin/listener.ora and changed the listening address to the new hostname:       (ADDRESS = (PROTOCOL = TCP)(HOST = vboxdb.oracle.com)(PORT = 1521)) After changing the listener.ora I was able to start the listener using: lsnrctl start I also had to reconfigure the EM database control.  I first deconfigured it using the command: emca -deconfig dbcontrol db -repos drop This drops the repository and removes any existing registered dbcontrols. I then re-configured it using the following command: emca -config dbcontrol db -repos create This creates the EM repository and then configures and starts dbcontrol. Now my database machine is ready so I can close it down and take a snapshot. Disabling the Database on the FMW Machine I set up the database to start automatically by creating a service called “dbora”.  On the FMW machine I do not need the database running so I can prevent it auto-starting by running the following command: chkconfig –del dbora Note that because I am using a snapshot it is not a waste of disk space to have the DB installed but not used.  As long as I don’t run it, it won’t cost me anything. I can now close the FMW machine down and take a snapshot. Creating a New Domain The FMW machine is now ready to create a new domain.  When creating the domain I can point it at the second machine which is running the database.  I can potentially run these machines on two separate physical machines as long as I have the original virtual machine available to both of the physical machines. Gotchas in Snapshotting VirtualBox does not support the concept of linked machines in a network like some virtualization technologies so when creating a snapshot it is a good idea to shut both VMs down and then take a snapshot on both of them.  This is because we want to keep the database in sync with the middleware.  One way to make sure that this happens would be to place all the domain configuration files on the database server via an NFS share, this would mean that all we would need to snapshot would be the database machine because that would hold all the state and configuration. The Sky’s the Limit We have covered a simple case of having just two machines.  I have a more complicated configuration in which two machine run a RAC database off the same base OS image, and two more machines run a SOA cluster based on the same OS image.  Just remember what machine holds state and what are the consequences of taking a snapshot.

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  • C#: String Concatenation vs Format vs StringBuilder

    - by James Michael Hare
    I was looking through my groups’ C# coding standards the other day and there were a couple of legacy items in there that caught my eye.  They had been passed down from committee to committee so many times that no one even thought to second guess and try them for a long time.  It’s yet another example of how micro-optimizations can often get the best of us and cause us to write code that is not as maintainable as it could be for the sake of squeezing an extra ounce of performance out of our software. So the two standards in question were these, in paraphrase: Prefer StringBuilder or string.Format() to string concatenation. Prefer string.Equals() with case-insensitive option to string.ToUpper().Equals(). Now some of you may already know what my results are going to show, as these items have been compared before on many blogs, but I think it’s always worth repeating and trying these yourself.  So let’s dig in. The first test was a pretty standard one.  When concattenating strings, what is the best choice: StringBuilder, string concattenation, or string.Format()? So before we being I read in a number of iterations from the console and a length of each string to generate.  Then I generate that many random strings of the given length and an array to hold the results.  Why am I so keen to keep the results?  Because I want to be able to snapshot the memory and don’t want garbage collection to collect the strings, hence the array to keep hold of them.  I also didn’t want the random strings to be part of the allocation, so I pre-allocate them and the array up front before the snapshot.  So in the code snippets below: num – Number of iterations. strings – Array of randomly generated strings. results – Array to hold the results of the concatenation tests. timer – A System.Diagnostics.Stopwatch() instance to time code execution. start – Beginning memory size. stop – Ending memory size. after – Memory size after final GC. So first, let’s look at the concatenation loop: 1: // build num strings using concattenation. 2: for (int i = 0; i < num; i++) 3: { 4: results[i] = "This is test #" + i + " with a result of " + strings[i]; 5: } Pretty standard, right?  Next for string.Format(): 1: // build strings using string.Format() 2: for (int i = 0; i < num; i++) 3: { 4: results[i] = string.Format("This is test #{0} with a result of {1}", i, strings[i]); 5: }   Finally, StringBuilder: 1: // build strings using StringBuilder 2: for (int i = 0; i < num; i++) 3: { 4: var builder = new StringBuilder(); 5: builder.Append("This is test #"); 6: builder.Append(i); 7: builder.Append(" with a result of "); 8: builder.Append(strings[i]); 9: results[i] = builder.ToString(); 10: } So I take each of these loops, and time them by using a block like this: 1: // get the total amount of memory used, true tells it to run GC first. 2: start = System.GC.GetTotalMemory(true); 3:  4: // restart the timer 5: timer.Reset(); 6: timer.Start(); 7:  8: // *** code to time and measure goes here. *** 9:  10: // get the current amount of memory, stop the timer, then get memory after GC. 11: stop = System.GC.GetTotalMemory(false); 12: timer.Stop(); 13: other = System.GC.GetTotalMemory(true); So let’s look at what happens when I run each of these blocks through the timer and memory check at 500,000 iterations: 1: Operator + - Time: 547, Memory: 56104540/55595960 - 500000 2: string.Format() - Time: 749, Memory: 57295812/55595960 - 500000 3: StringBuilder - Time: 608, Memory: 55312888/55595960 – 500000   Egad!  string.Format brings up the rear and + triumphs, well, at least in terms of speed.  The concat burns more memory than StringBuilder but less than string.Format().  This shows two main things: StringBuilder is not always the panacea many think it is. The difference between any of the three is miniscule! The second point is extremely important!  You will often here people who will grasp at results and say, “look, operator + is 10% faster than StringBuilder so always use StringBuilder.”  Statements like this are a disservice and often misleading.  For example, if I had a good guess at what the size of the string would be, I could have preallocated my StringBuffer like so:   1: for (int i = 0; i < num; i++) 2: { 3: // pre-declare StringBuilder to have 100 char buffer. 4: var builder = new StringBuilder(100); 5: builder.Append("This is test #"); 6: builder.Append(i); 7: builder.Append(" with a result of "); 8: builder.Append(strings[i]); 9: results[i] = builder.ToString(); 10: }   Now let’s look at the times: 1: Operator + - Time: 551, Memory: 56104412/55595960 - 500000 2: string.Format() - Time: 753, Memory: 57296484/55595960 - 500000 3: StringBuilder - Time: 525, Memory: 59779156/55595960 - 500000   Whoa!  All of the sudden StringBuilder is back on top again!  But notice, it takes more memory now.  This makes perfect sense if you examine the IL behind the scenes.  Whenever you do a string concat (+) in your code, it examines the lengths of the arguments and creates a StringBuilder behind the scenes of the appropriate size for you. But even IF we know the approximate size of our StringBuilder, look how much less readable it is!  That’s why I feel you should always take into account both readability and performance.  After all, consider all these timings are over 500,000 iterations.   That’s at best  0.0004 ms difference per call which is neglidgable at best.  The key is to pick the best tool for the job.  What do I mean?  Consider these awesome words of wisdom: Concatenate (+) is best at concatenating.  StringBuilder is best when you need to building. Format is best at formatting. Totally Earth-shattering, right!  But if you consider it carefully, it actually has a lot of beauty in it’s simplicity.  Remember, there is no magic bullet.  If one of these always beat the others we’d only have one and not three choices. The fact is, the concattenation operator (+) has been optimized for speed and looks the cleanest for joining together a known set of strings in the simplest manner possible. StringBuilder, on the other hand, excels when you need to build a string of inderterminant length.  Use it in those times when you are looping till you hit a stop condition and building a result and it won’t steer you wrong. String.Format seems to be the looser from the stats, but consider which of these is more readable.  Yes, ignore the fact that you could do this with ToString() on a DateTime.  1: // build a date via concatenation 2: var date1 = (month < 10 ? string.Empty : "0") + month + '/' 3: + (day < 10 ? string.Empty : "0") + '/' + year; 4:  5: // build a date via string builder 6: var builder = new StringBuilder(10); 7: if (month < 10) builder.Append('0'); 8: builder.Append(month); 9: builder.Append('/'); 10: if (day < 10) builder.Append('0'); 11: builder.Append(day); 12: builder.Append('/'); 13: builder.Append(year); 14: var date2 = builder.ToString(); 15:  16: // build a date via string.Format 17: var date3 = string.Format("{0:00}/{1:00}/{2:0000}", month, day, year); 18:  So the strength in string.Format is that it makes constructing a formatted string easy to read.  Yes, it’s slower, but look at how much more elegant it is to do zero-padding and anything else string.Format does. So my lesson is, don’t look for the silver bullet!  Choose the best tool.  Micro-optimization almost always bites you in the end because you’re sacrificing readability for performance, which is almost exactly the wrong choice 90% of the time. I love the rules of optimization.  They’ve been stated before in many forms, but here’s how I always remember them: For Beginners: Do not optimize. For Experts: Do not optimize yet. It’s so true.  Most of the time on today’s modern hardware, a micro-second optimization at the sake of readability will net you nothing because it won’t be your bottleneck.  Code for readability, choose the best tool for the job which will usually be the most readable and maintainable as well.  Then, and only then, if you need that extra performance boost after profiling your code and exhausting all other options… then you can start to think about optimizing.

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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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  • Hello Operator, My Switch Is Bored

    - by Paul White
    This is a post for T-SQL Tuesday #43 hosted by my good friend Rob Farley. The topic this month is Plan Operators. I haven’t taken part in T-SQL Tuesday before, but I do like to write about execution plans, so this seemed like a good time to start. This post is in two parts. The first part is primarily an excuse to use a pretty bad play on words in the title of this blog post (if you’re too young to know what a telephone operator or a switchboard is, I hate you). The second part of the post looks at an invisible query plan operator (so to speak). 1. My Switch Is Bored Allow me to present the rare and interesting execution plan operator, Switch: Books Online has this to say about Switch: Following that description, I had a go at producing a Fast Forward Cursor plan that used the TOP operator, but had no luck. That may be due to my lack of skill with cursors, I’m not too sure. The only application of Switch in SQL Server 2012 that I am familiar with requires a local partitioned view: CREATE TABLE dbo.T1 (c1 int NOT NULL CHECK (c1 BETWEEN 00 AND 24)); CREATE TABLE dbo.T2 (c1 int NOT NULL CHECK (c1 BETWEEN 25 AND 49)); CREATE TABLE dbo.T3 (c1 int NOT NULL CHECK (c1 BETWEEN 50 AND 74)); CREATE TABLE dbo.T4 (c1 int NOT NULL CHECK (c1 BETWEEN 75 AND 99)); GO CREATE VIEW V1 AS SELECT c1 FROM dbo.T1 UNION ALL SELECT c1 FROM dbo.T2 UNION ALL SELECT c1 FROM dbo.T3 UNION ALL SELECT c1 FROM dbo.T4; Not only that, but it needs an updatable local partitioned view. We’ll need some primary keys to meet that requirement: ALTER TABLE dbo.T1 ADD CONSTRAINT PK_T1 PRIMARY KEY (c1);   ALTER TABLE dbo.T2 ADD CONSTRAINT PK_T2 PRIMARY KEY (c1);   ALTER TABLE dbo.T3 ADD CONSTRAINT PK_T3 PRIMARY KEY (c1);   ALTER TABLE dbo.T4 ADD CONSTRAINT PK_T4 PRIMARY KEY (c1); We also need an INSERT statement that references the view. Even more specifically, to see a Switch operator, we need to perform a single-row insert (multi-row inserts use a different plan shape): INSERT dbo.V1 (c1) VALUES (1); And now…the execution plan: The Constant Scan manufactures a single row with no columns. The Compute Scalar works out which partition of the view the new value should go in. The Assert checks that the computed partition number is not null (if it is, an error is returned). The Nested Loops Join executes exactly once, with the partition id as an outer reference (correlated parameter). The Switch operator checks the value of the parameter and executes the corresponding input only. If the partition id is 0, the uppermost Clustered Index Insert is executed, adding a row to table T1. If the partition id is 1, the next lower Clustered Index Insert is executed, adding a row to table T2…and so on. In case you were wondering, here’s a query and execution plan for a multi-row insert to the view: INSERT dbo.V1 (c1) VALUES (1), (2); Yuck! An Eager Table Spool and four Filters! I prefer the Switch plan. My guess is that almost all the old strategies that used a Switch operator have been replaced over time, using things like a regular Concatenation Union All combined with Start-Up Filters on its inputs. Other new (relative to the Switch operator) features like table partitioning have specific execution plan support that doesn’t need the Switch operator either. This feels like a bit of a shame, but perhaps it is just nostalgia on my part, it’s hard to know. Please do let me know if you encounter a query that can still use the Switch operator in 2012 – it must be very bored if this is the only possible modern usage! 2. Invisible Plan Operators The second part of this post uses an example based on a question Dave Ballantyne asked using the SQL Sentry Plan Explorer plan upload facility. If you haven’t tried that yet, make sure you’re on the latest version of the (free) Plan Explorer software, and then click the Post to SQLPerformance.com button. That will create a site question with the query plan attached (which can be anonymized if the plan contains sensitive information). Aaron Bertrand and I keep a close eye on questions there, so if you have ever wanted to ask a query plan question of either of us, that’s a good way to do it. The problem The issue I want to talk about revolves around a query issued against a calendar table. The script below creates a simplified version and adds 100 years of per-day information to it: USE tempdb; GO CREATE TABLE dbo.Calendar ( dt date NOT NULL, isWeekday bit NOT NULL, theYear smallint NOT NULL,   CONSTRAINT PK__dbo_Calendar_dt PRIMARY KEY CLUSTERED (dt) ); GO -- Monday is the first day of the week for me SET DATEFIRST 1;   -- Add 100 years of data INSERT dbo.Calendar WITH (TABLOCKX) (dt, isWeekday, theYear) SELECT CA.dt, isWeekday = CASE WHEN DATEPART(WEEKDAY, CA.dt) IN (6, 7) THEN 0 ELSE 1 END, theYear = YEAR(CA.dt) FROM Sandpit.dbo.Numbers AS N CROSS APPLY ( VALUES (DATEADD(DAY, N.n - 1, CONVERT(date, '01 Jan 2000', 113))) ) AS CA (dt) WHERE N.n BETWEEN 1 AND 36525; The following query counts the number of weekend days in 2013: SELECT Days = COUNT_BIG(*) FROM dbo.Calendar AS C WHERE theYear = 2013 AND isWeekday = 0; It returns the correct result (104) using the following execution plan: The query optimizer has managed to estimate the number of rows returned from the table exactly, based purely on the default statistics created separately on the two columns referenced in the query’s WHERE clause. (Well, almost exactly, the unrounded estimate is 104.289 rows.) There is already an invisible operator in this query plan – a Filter operator used to apply the WHERE clause predicates. We can see it by re-running the query with the enormously useful (but undocumented) trace flag 9130 enabled: Now we can see the full picture. The whole table is scanned, returning all 36,525 rows, before the Filter narrows that down to just the 104 we want. Without the trace flag, the Filter is incorporated in the Clustered Index Scan as a residual predicate. It is a little bit more efficient than using a separate operator, but residual predicates are still something you will want to avoid where possible. The estimates are still spot on though: Anyway, looking to improve the performance of this query, Dave added the following filtered index to the Calendar table: CREATE NONCLUSTERED INDEX Weekends ON dbo.Calendar(theYear) WHERE isWeekday = 0; The original query now produces a much more efficient plan: Unfortunately, the estimated number of rows produced by the seek is now wrong (365 instead of 104): What’s going on? The estimate was spot on before we added the index! Explanation You might want to grab a coffee for this bit. Using another trace flag or two (8606 and 8612) we can see that the cardinality estimates were exactly right initially: The highlighted information shows the initial cardinality estimates for the base table (36,525 rows), the result of applying the two relational selects in our WHERE clause (104 rows), and after performing the COUNT_BIG(*) group by aggregate (1 row). All of these are correct, but that was before cost-based optimization got involved :) Cost-based optimization When cost-based optimization starts up, the logical tree above is copied into a structure (the ‘memo’) that has one group per logical operation (roughly speaking). The logical read of the base table (LogOp_Get) ends up in group 7; the two predicates (LogOp_Select) end up in group 8 (with the details of the selections in subgroups 0-6). These two groups still have the correct cardinalities as trace flag 8608 output (initial memo contents) shows: During cost-based optimization, a rule called SelToIdxStrategy runs on group 8. It’s job is to match logical selections to indexable expressions (SARGs). It successfully matches the selections (theYear = 2013, is Weekday = 0) to the filtered index, and writes a new alternative into the memo structure. The new alternative is entered into group 8 as option 1 (option 0 was the original LogOp_Select): The new alternative is to do nothing (PhyOp_NOP = no operation), but to instead follow the new logical instructions listed below the NOP. The LogOp_GetIdx (full read of an index) goes into group 21, and the LogOp_SelectIdx (selection on an index) is placed in group 22, operating on the result of group 21. The definition of the comparison ‘the Year = 2013’ (ScaOp_Comp downwards) was already present in the memo starting at group 2, so no new memo groups are created for that. New Cardinality Estimates The new memo groups require two new cardinality estimates to be derived. First, LogOp_Idx (full read of the index) gets a predicted cardinality of 10,436. This number comes from the filtered index statistics: DBCC SHOW_STATISTICS (Calendar, Weekends) WITH STAT_HEADER; The second new cardinality derivation is for the LogOp_SelectIdx applying the predicate (theYear = 2013). To get a number for this, the cardinality estimator uses statistics for the column ‘theYear’, producing an estimate of 365 rows (there are 365 days in 2013!): DBCC SHOW_STATISTICS (Calendar, theYear) WITH HISTOGRAM; This is where the mistake happens. Cardinality estimation should have used the filtered index statistics here, to get an estimate of 104 rows: DBCC SHOW_STATISTICS (Calendar, Weekends) WITH HISTOGRAM; Unfortunately, the logic has lost sight of the link between the read of the filtered index (LogOp_GetIdx) in group 22, and the selection on that index (LogOp_SelectIdx) that it is deriving a cardinality estimate for, in group 21. The correct cardinality estimate (104 rows) is still present in the memo, attached to group 8, but that group now has a PhyOp_NOP implementation. Skipping over the rest of cost-based optimization (in a belated attempt at brevity) we can see the optimizer’s final output using trace flag 8607: This output shows the (incorrect, but understandable) 365 row estimate for the index range operation, and the correct 104 estimate still attached to its PhyOp_NOP. This tree still has to go through a few post-optimizer rewrites and ‘copy out’ from the memo structure into a tree suitable for the execution engine. One step in this process removes PhyOp_NOP, discarding its 104-row cardinality estimate as it does so. To finish this section on a more positive note, consider what happens if we add an OVER clause to the query aggregate. This isn’t intended to be a ‘fix’ of any sort, I just want to show you that the 104 estimate can survive and be used if later cardinality estimation needs it: SELECT Days = COUNT_BIG(*) OVER () FROM dbo.Calendar AS C WHERE theYear = 2013 AND isWeekday = 0; The estimated execution plan is: Note the 365 estimate at the Index Seek, but the 104 lives again at the Segment! We can imagine the lost predicate ‘isWeekday = 0’ as sitting between the seek and the segment in an invisible Filter operator that drops the estimate from 365 to 104. Even though the NOP group is removed after optimization (so we don’t see it in the execution plan) bear in mind that all cost-based choices were made with the 104-row memo group present, so although things look a bit odd, it shouldn’t affect the optimizer’s plan selection. I should also mention that we can work around the estimation issue by including the index’s filtering columns in the index key: CREATE NONCLUSTERED INDEX Weekends ON dbo.Calendar(theYear, isWeekday) WHERE isWeekday = 0 WITH (DROP_EXISTING = ON); There are some downsides to doing this, including that changes to the isWeekday column may now require Halloween Protection, but that is unlikely to be a big problem for a static calendar table ;)  With the updated index in place, the original query produces an execution plan with the correct cardinality estimation showing at the Index Seek: That’s all for today, remember to let me know about any Switch plans you come across on a modern instance of SQL Server! Finally, here are some other posts of mine that cover other plan operators: Segment and Sequence Project Common Subexpression Spools Why Plan Operators Run Backwards Row Goals and the Top Operator Hash Match Flow Distinct Top N Sort Index Spools and Page Splits Singleton and Range Seeks Bitmaps Hash Join Performance Compute Scalar © 2013 Paul White – All Rights Reserved Twitter: @SQL_Kiwi

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  • VLOOKUP in Excel, part 2: Using VLOOKUP without a database

    - by Mark Virtue
    In a recent article, we introduced the Excel function called VLOOKUP and explained how it could be used to retrieve information from a database into a cell in a local worksheet.  In that article we mentioned that there were two uses for VLOOKUP, and only one of them dealt with querying databases.  In this article, the second and final in the VLOOKUP series, we examine this other, lesser known use for the VLOOKUP function. If you haven’t already done so, please read the first VLOOKUP article – this article will assume that many of the concepts explained in that article are already known to the reader. When working with databases, VLOOKUP is passed a “unique identifier” that serves to identify which data record we wish to find in the database (e.g. a product code or customer ID).  This unique identifier must exist in the database, otherwise VLOOKUP returns us an error.  In this article, we will examine a way of using VLOOKUP where the identifier doesn’t need to exist in the database at all.  It’s almost as if VLOOKUP can adopt a “near enough is good enough” approach to returning the data we’re looking for.  In certain circumstances, this is exactly what we need. We will illustrate this article with a real-world example – that of calculating the commissions that are generated on a set of sales figures.  We will start with a very simple scenario, and then progressively make it more complex, until the only rational solution to the problem is to use VLOOKUP.  The initial scenario in our fictitious company works like this:  If a salesperson creates more than $30,000 worth of sales in a given year, the commission they earn on those sales is 30%.  Otherwise their commission is only 20%.  So far this is a pretty simple worksheet: To use this worksheet, the salesperson enters their sales figures in cell B1, and the formula in cell B2 calculates the correct commission rate they are entitled to receive, which is used in cell B3 to calculate the total commission that the salesperson is owed (which is a simple multiplication of B1 and B2). The cell B2 contains the only interesting part of this worksheet – the formula for deciding which commission rate to use: the one below the threshold of $30,000, or the one above the threshold.  This formula makes use of the Excel function called IF.  For those readers that are not familiar with IF, it works like this: IF(condition,value if true,value if false) Where the condition is an expression that evaluates to either true or false.  In the example above, the condition is the expression B1<B5, which can be read as “Is B1 less than B5?”, or, put another way, “Are the total sales less than the threshold”.  If the answer to this question is “yes” (true), then we use the value if true parameter of the function, namely B6 in this case – the commission rate if the sales total was below the threshold.  If the answer to the question is “no” (false), then we use the value if false parameter of the function, namely B7 in this case – the commission rate if the sales total was above the threshold. As you can see, using a sales total of $20,000 gives us a commission rate of 20% in cell B2.  If we enter a value of $40,000, we get a different commission rate: So our spreadsheet is working. Let’s make it more complex.  Let’s introduce a second threshold:  If the salesperson earns more than $40,000, then their commission rate increases to 40%: Easy enough to understand in the real world, but in cell B2 our formula is getting more complex.  If you look closely at the formula, you’ll see that the third parameter of the original IF function (the value if false) is now an entire IF function in its own right.  This is called a nested function (a function within a function).  It’s perfectly valid in Excel (it even works!), but it’s harder to read and understand. We’re not going to go into the nuts and bolts of how and why this works, nor will we examine the nuances of nested functions.  This is a tutorial on VLOOKUP, not on Excel in general. Anyway, it gets worse!  What about when we decide that if they earn more than $50,000 then they’re entitled to 50% commission, and if they earn more than $60,000 then they’re entitled to 60% commission? Now the formula in cell B2, while correct, has become virtually unreadable.  No-one should have to write formulae where the functions are nested four levels deep!  Surely there must be a simpler way? There certainly is.  VLOOKUP to the rescue! Let’s redesign the worksheet a bit.  We’ll keep all the same figures, but organize it in a new way, a more tabular way: Take a moment and verify for yourself that the new Rate Table works exactly the same as the series of thresholds above. Conceptually, what we’re about to do is use VLOOKUP to look up the salesperson’s sales total (from B1) in the rate table and return to us the corresponding commission rate.  Note that the salesperson may have indeed created sales that are not one of the five values in the rate table ($0, $30,000, $40,000, $50,000 or $60,000).  They may have created sales of $34,988.  It’s important to note that $34,988 does not appear in the rate table.  Let’s see if VLOOKUP can solve our problem anyway… We select cell B2 (the location we want to put our formula), and then insert the VLOOKUP function from the Formulas tab: The Function Arguments box for VLOOKUP appears.  We fill in the arguments (parameters) one by one, starting with the Lookup_value, which is, in this case, the sales total from cell B1.  We place the cursor in the Lookup_value field and then click once on cell B1: Next we need to specify to VLOOKUP what table to lookup this data in.  In this example, it’s the rate table, of course.  We place the cursor in the Table_array field, and then highlight the entire rate table – excluding the headings: Next we must specify which column in the table contains the information we want our formula to return to us.  In this case we want the commission rate, which is found in the second column in the table, so we therefore enter a 2 into the Col_index_num field: Finally we enter a value in the Range_lookup field. Important:  It is the use of this field that differentiates the two ways of using VLOOKUP.  To use VLOOKUP with a database, this final parameter, Range_lookup, must always be set to FALSE, but with this other use of VLOOKUP, we must either leave it blank or enter a value of TRUE.  When using VLOOKUP, it is vital that you make the correct choice for this final parameter. To be explicit, we will enter a value of true in the Range_lookup field.  It would also be fine to leave it blank, as this is the default value: We have completed all the parameters.  We now click the OK button, and Excel builds our VLOOKUP formula for us: If we experiment with a few different sales total amounts, we can satisfy ourselves that the formula is working. Conclusion In the “database” version of VLOOKUP, where the Range_lookup parameter is FALSE, the value passed in the first parameter (Lookup_value) must be present in the database.  In other words, we’re looking for an exact match. But in this other use of VLOOKUP, we are not necessarily looking for an exact match.  In this case, “near enough is good enough”.  But what do we mean by “near enough”?  Let’s use an example:  When searching for a commission rate on a sales total of $34,988, our VLOOKUP formula will return us a value of 30%, which is the correct answer.  Why did it choose the row in the table containing 30% ?  What, in fact, does “near enough” mean in this case?  Let’s be precise: When Range_lookup is set to TRUE (or omitted), VLOOKUP will look in column 1 and match the highest value that is not greater than the Lookup_value parameter. It’s also important to note that for this system to work, the table must be sorted in ascending order on column 1! If you would like to practice with VLOOKUP, the sample file illustrated in this article can be downloaded from here. Similar Articles Productive Geek Tips Using VLOOKUP in ExcelImport Microsoft Access Data Into ExcelImport an Access Database into ExcelCopy a Group of Cells in Excel 2007 to the Clipboard as an ImageShare Access Data with Excel in Office 2010 TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Quickly Schedule Meetings With NeedtoMeet Share Flickr Photos On Facebook Automatically Are You Blocked On Gtalk? Find out Discover Latest Android Apps On AppBrain The Ultimate Guide For YouTube Lovers Will it Blend? iPad Edition

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  • Caching NHibernate Named Queries

    - by TStewartDev
    I recently started a new job and one of my first tasks was to implement a "popular products" design. The parameters were that it be done with NHibernate and be cached for 24 hours at a time because the query will be pretty taxing and the results do not need to be constantly up to date. This ended up being tougher than it sounds. The database schema meant a minimum of four joins with filtering and ordering criteria. I decided to use a stored procedure rather than letting NHibernate create the SQL for me. Here is a summary of what I learned (even if I didn't ultimately use all of it): You can't, at the time of this writing, use Fluent NHibernate to configure SQL named queries or imports You can return persistent entities from a stored procedure and there are a couple ways to do that You can populate POCOs using the results of a stored procedure, but it isn't quite as obvious You can reuse your named query result mapping other places (avoid duplication) Caching your query results is not at all obvious Testing to see if your cache is working is a pain NHibernate does a lot of things right. Having unified, up-to-date, comprehensive, and easy-to-find documentation is not one of them. By the way, if you're new to this, I'll use the terms "named query" and "stored procedure" (from NHibernate's perspective) fairly interchangeably. Technically, a named query can execute any SQL, not just a stored procedure, and a stored procedure doesn't have to be executed from a named query, but for reusability, it seems to me like the best practice. If you're here, chances are good you're looking for answers to a similar problem. You don't want to read about the path, you just want the result. So, here's how to get this thing going. The Stored Procedure NHibernate has some guidelines when using stored procedures. For Microsoft SQL Server, you have to return a result set. The scalar value that the stored procedure returns is ignored as are any result sets after the first. Other than that, it's nothing special. CREATE PROCEDURE GetPopularProducts @StartDate DATETIME, @MaxResults INT AS BEGIN SELECT [ProductId], [ProductName], [ImageUrl] FROM SomeTableWithJoinsEtc END The Result Class - PopularProduct You have two options to transport your query results to your view (or wherever is the final destination): you can populate an existing mapped entity class in your model, or you can create a new entity class. If you go with the existing model, the advantage is that the query will act as a loader and you'll get full proxied access to the domain model. However, this can be a disadvantage if you require access to the related entities that aren't loaded by your results. For example, my PopularProduct has image references. Unless I tie them into the query (thus making it even more complicated and expensive to run), they'll have to be loaded on access, requiring more trips to the database. Since we're trying to avoid trips to the database by using a second-level cache, we should use the second option, which is to create a separate entity for results. This approach is (I believe) in the spirit of the Command-Query Separation principle, and it allows us to flatten our data and optimize our report-generation process from data source to view. public class PopularProduct { public virtual int ProductId { get; set; } public virtual string ProductName { get; set; } public virtual string ImageUrl { get; set; } } The NHibernate Mappings (hbm) Next up, we need to let NHibernate know about the query and where the results will go. Below is the markup for the PopularProduct class. Notice that I'm using the <resultset> element and that it has a name attribute. The name allows us to drop this into our query map and any others, giving us reusability. Also notice the <import> element which lets NHibernate know about our entity class. <?xml version="1.0" encoding="utf-8" ?> <hibernate-mapping xmlns="urn:nhibernate-mapping-2.2"> <import class="PopularProduct, Infrastructure.NHibernate, Version=1.0.0.0"/> <resultset name="PopularProductResultSet"> <return-scalar column="ProductId" type="System.Int32"/> <return-scalar column="ProductName" type="System.String"/> <return-scalar column="ImageUrl" type="System.String"/> </resultset> </hibernate-mapping>  And now the PopularProductsMap: <?xml version="1.0" encoding="utf-8" ?> <hibernate-mapping xmlns="urn:nhibernate-mapping-2.2"> <sql-query name="GetPopularProducts" resultset-ref="PopularProductResultSet" cacheable="true" cache-mode="normal"> <query-param name="StartDate" type="System.DateTime" /> <query-param name="MaxResults" type="System.Int32" /> exec GetPopularProducts @StartDate = :StartDate, @MaxResults = :MaxResults </sql-query> </hibernate-mapping>  The two most important things to notice here are the resultset-ref attribute, which links in our resultset mapping, and the cacheable attribute. The Query Class – PopularProductsQuery So far, this has been fairly obvious if you're familiar with NHibernate. This next part, maybe not so much. You can implement your query however you want to; for me, I wanted a self-encapsulated Query class, so here's what it looks like: public class PopularProductsQuery : IPopularProductsQuery { private static readonly IResultTransformer ResultTransformer; private readonly ISessionBuilder _sessionBuilder;   static PopularProductsQuery() { ResultTransformer = Transformers.AliasToBean<PopularProduct>(); }   public PopularProductsQuery(ISessionBuilder sessionBuilder) { _sessionBuilder = sessionBuilder; }   public IList<PopularProduct> GetPopularProducts(DateTime startDate, int maxResults) { var session = _sessionBuilder.GetSession(); var popularProducts = session .GetNamedQuery("GetPopularProducts") .SetCacheable(true) .SetCacheRegion("PopularProductsCacheRegion") .SetCacheMode(CacheMode.Normal) .SetReadOnly(true) .SetResultTransformer(ResultTransformer) .SetParameter("StartDate", startDate.Date) .SetParameter("MaxResults", maxResults) .List<PopularProduct>();   return popularProducts; } }  Okay, so let's look at each line of the query execution. The first, GetNamedQuery, matches up with our NHibernate mapping for the sql-query. Next, we set it as cacheable (this is probably redundant since our mapping also specified it, but it can't hurt, right?). Then we set the cache region which we'll get to in the next section. Set the cache mode (optional, I believe), and my cache is read-only, so I set that as well. The result transformer is very important. This tells NHibernate how to transform your query results into a non-persistent entity. You can see I've defined ResultTransformer in the static constructor using the AliasToBean transformer. The name is obviously leftover from Java/Hibernate. Finally, set your parameters and then call a result method which will execute the query. Because this is set to cached, you execute this statement every time you run the query and NHibernate will know based on your parameters whether to use its cached version or a fresh version. The Configuration – hibernate.cfg.xml and Web.config You need to explicitly enable second-level caching in your hibernate configuration: <hibernate-configuration xmlns="urn:nhibernate-configuration-2.2"> <session-factory> [...] <property name="dialect">NHibernate.Dialect.MsSql2005Dialect</property> <property name="cache.provider_class">NHibernate.Caches.SysCache.SysCacheProvider,NHibernate.Caches.SysCache</property> <property name="cache.use_query_cache">true</property> <property name="cache.use_second_level_cache">true</property> [...] </session-factory> </hibernate-configuration> Both properties "use_query_cache" and "use_second_level_cache" are necessary. As this is for a web deployement, we're using SysCache which relies on ASP.NET's caching. Be aware of this if you're not deploying to the web! You'll have to use a different cache provider. We also need to tell our cache provider (in this cache, SysCache) about our caching region: <syscache> <cache region="PopularProductsCacheRegion" expiration="86400" priority="5" /> </syscache> Here I've set the cache to be valid for 24 hours. This XML snippet goes in your Web.config (or in a separate file referenced by Web.config, which helps keep things tidy). The Payoff That should be it! At this point, your queries should run once against the database for a given set of parameters and then use the cache thereafter until it expires. You can, of course, adjust settings to work in your particular environment. Testing Testing your application to ensure it is using the cache is a pain, but if you're like me, you want to know that it's actually working. It's a bit involved, though, so I'll create a separate post for it if comments indicate there is interest.

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  • How to tell if SPARC T4 crypto is being used?

    - by danx
    A question that often comes up when running applications on SPARC T4 systems is "How can I tell if hardware crypto accleration is being used?" To review, the SPARC T4 processor includes a crypto unit that supports several crypto instructions. For hardware crypto these include 11 AES instructions, 4 xmul* instructions (for AES GCM carryless multiply), mont for Montgomery multiply (optimizes RSA and DSA), and 5 des_* instructions (for DES3). For hardware hash algorithm optimization, the T4 has the md5, sha1, sha256, and sha512 instructions (the last two are used for SHA-224 an SHA-384). First off, it's easy to tell if the processor T4 crypto instructions—use the isainfo -v command and look for "sparcv9" and "aes" (and other hash and crypto algorithms) in the output: $ isainfo -v 64-bit sparcv9 applications crc32c cbcond pause mont mpmul sha512 sha256 sha1 md5 camellia kasumi des aes ima hpc vis3 fmaf asi_blk_init vis2 vis popc These instructions are not-privileged, so are available for direct use in user-level applications and libraries (such as OpenSSL). Here is the "openssl speed -evp" command shown with the built-in t4 engine and with the pkcs11 engine. Both run the T4 AES instructions, but the t4 engine is faster than the pkcs11 engine because it has less overhead (especially for smaller packet sizes): t-4 $ /usr/bin/openssl version OpenSSL 1.0.0j 10 May 2012 t-4 $ /usr/bin/openssl engine (t4) SPARC T4 engine support (dynamic) Dynamic engine loading support (pkcs11) PKCS #11 engine support t-4 $ /usr/bin/openssl speed -evp aes-128-cbc # t4 engine used by default . . . The 'numbers' are in 1000s of bytes per second processed. type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 487777.10k 816822.21k 986012.59k 1017029.97k 1053543.08k t-4 $ /usr/bin/openssl speed -engine pkcs11 -evp aes-128-cbc engine "pkcs11" set. . . . The 'numbers' are in 1000s of bytes per second processed. type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 31703.58k 116636.39k 350672.81k 696170.50k 993599.49k Note: The "-evp" flag indicates use the OpenSSL "EnVeloPe" API, which gives more accurate results. That's because it tells OpenSSL to use the same API that external programs use when calling OpenSSL libcrypto functions, evp(3openssl). DTrace Shows if T4 Crypto Functions Are Used OK, good enough, the isainfo(1) command shows the instructions are present, but how does one know if they are being used? Chi-Chang Lin, who works on Oracle Solaris performance, wrote a Dtrace script to show if T4 instructions are being executed. To show the T4 instructions are being used, run the following Dtrace script. Look for functions named "t4" and "yf" in the output. The OpenSSL T4 engine uses functions named "t4" and the PKCS#11 engine uses functions named "yf". To demonstrate, I'll first run "openssl speed" with the built-in t4 engine then with the pkcs11 engine. The performance numbers are not valid due to dtrace probes slowing things down. t-4 # dtrace -Z -n ' pid$target::*yf*:entry,pid$target::*t4_*:entry{ @[probemod, probefunc] = count();}' \ -c "/usr/bin/openssl speed -evp aes-128-cbc" dtrace: description 'pid$target::*yf*:entry' matched 101 probes . . . dtrace: pid 2029 has exited libcrypto.so.1.0.0 ENGINE_load_t4 1 libcrypto.so.1.0.0 t4_DH 1 libcrypto.so.1.0.0 t4_DSA 1 libcrypto.so.1.0.0 t4_RSA 1 libcrypto.so.1.0.0 t4_destroy 1 libcrypto.so.1.0.0 t4_free_aes_ctr_NIDs 1 libcrypto.so.1.0.0 t4_init 1 libcrypto.so.1.0.0 t4_add_NID 3 libcrypto.so.1.0.0 t4_aes_expand128 5 libcrypto.so.1.0.0 t4_cipher_init_aes 5 libcrypto.so.1.0.0 t4_get_all_ciphers 6 libcrypto.so.1.0.0 t4_get_all_digests 59 libcrypto.so.1.0.0 t4_digest_final_sha1 65 libcrypto.so.1.0.0 t4_digest_init_sha1 65 libcrypto.so.1.0.0 t4_sha1_multiblock 126 libcrypto.so.1.0.0 t4_digest_update_sha1 261 libcrypto.so.1.0.0 t4_aes128_cbc_encrypt 1432979 libcrypto.so.1.0.0 t4_aes128_load_keys_for_encrypt 1432979 libcrypto.so.1.0.0 t4_cipher_do_aes_128_cbc 1432979 t-4 # dtrace -Z -n 'pid$target::*yf*:entry{ @[probemod, probefunc] = count();}   pid$target::*yf*:entry,pid$target::*t4_*:entry{ @[probemod, probefunc] = count();}' \ -c "/usr/bin/openssl speed -engine pkcs11 -evp aes-128-cbc" dtrace: description 'pid$target::*yf*:entry' matched 101 probes engine "pkcs11" set. . . . dtrace: pid 2033 has exited libcrypto.so.1.0.0 ENGINE_load_t4 1 libcrypto.so.1.0.0 t4_DH 1 libcrypto.so.1.0.0 t4_DSA 1 libcrypto.so.1.0.0 t4_RSA 1 libcrypto.so.1.0.0 t4_destroy 1 libcrypto.so.1.0.0 t4_free_aes_ctr_NIDs 1 libcrypto.so.1.0.0 t4_get_all_ciphers 1 libcrypto.so.1.0.0 t4_get_all_digests 1 libsoftcrypto.so.1 rijndael_key_setup_enc_yf 1 libsoftcrypto.so.1 yf_aes_expand128 1 libcrypto.so.1.0.0 t4_add_NID 3 libsoftcrypto.so.1 yf_aes128_cbc_encrypt 1542330 libsoftcrypto.so.1 yf_aes128_load_keys_for_encrypt 1542330 So, as shown above the OpenSSL built-in t4 engine executes t4_* functions (which are hand-coded assembly executing the T4 AES instructions) and the OpenSSL pkcs11 engine executes *yf* functions. Programmatic Use of OpenSSL T4 engine The OpenSSL t4 engine is used automatically with the /usr/bin/openssl command line. Chi-Chang Lin also points out that if you're calling the OpenSSL API (libcrypto.so) from a program, you must call ENGINE_load_built_engines(), otherwise the built-in t4 engine will not be loaded. You do not call ENGINE_set_default(). That's because "openssl speed -evp" test calls ENGINE_load_built_engines() even though the "-engine" option wasn't specified. OpenSSL T4 engine Availability The OpenSSL t4 engine is available with Solaris 11 and 11.1. For Solaris 10 08/11 (U10), you need to use the OpenSSL pkcs311 engine. The OpenSSL t4 engine is distributed only with the version of OpenSSL distributed with Solaris (and not third-party or self-compiled versions of OpenSSL). The OpenSSL engine implements the AES cipher for Solaris 11, released 11/2011. For Solaris 11.1, released 11/2012, the OpenSSL engine adds optimization for the MD5, SHA-1, and SHA-2 hash algorithms, and DES-3. Although the T4 processor has Camillia and Kasumi block cipher instructions, these are not implemented in the OpenSSL T4 engine. The following charts may help view availability of optimizations. The first chart shows what's available with Solaris CLIs and APIs, the second chart shows what's available in Solaris OpenSSL. Native Solaris Optimization for SPARC T4 This table is shows Solaris native CLI and API support. As such, they are all available with the OpenSSL pkcs11 engine. CLIs: "openssl -engine pkcs11", encrypt(1), decrypt(1), mac(1), digest(1), MD5sum(1), SHA1sum(1), SHA224sum(1), SHA256sum(1), SHA384sum(1), SHA512sum(1) APIs: PKCS#11 library libpkcs11(3LIB) (incluDES Openssl pkcs11 engine), libMD(3LIB), and Solaris kernel modules AlgorithmSolaris 1008/11 (U10)Solaris 11Solaris 11.1 AES-ECB, AES-CBC, AES-CTR, AES-CBC AES-CFB128 XXX DES3-ECB, DES3-CBC, DES2-ECB, DES2-CBC, DES-ECB, DES-CBC XXX bignum Montgomery multiply (RSA, DSA) XXX MD5, SHA-1, SHA-256, SHA-384, SHA-512 XXX SHA-224 X ARCFOUR (RC4) X Solaris OpenSSL T4 Engine Optimization This table is for the Solaris OpenSSL built-in t4 engine. Algorithms listed above are also available through the OpenSSL pkcs11 engine. CLI: openssl(1openssl) APIs: openssl(5), engine(3openssl), evp(3openssl), libcrypto crypto(3openssl) AlgorithmSolaris 11Solaris 11SRU2Solaris 11.1 AES-ECB, AES-CBC, AES-CTR, AES-CBC AES-CFB128 XXX DES3-ECB, DES3-CBC, DES-ECB, DES-CBC X bignum Montgomery multiply (RSA, DSA) X MD5, SHA-1, SHA-256, SHA-384, SHA-512 XX SHA-224 X Source Code Availability Solaris Most of the T4 assembly code that called the new T4 crypto instructions was written by Ferenc Rákóczi of the Solaris Security group, with assistance from others. You can download the Solaris source for this and other parts of Solaris as a few zip files at the Oracle Download website. The relevant source files are generally under directories usr/src/common/crypto/{aes,arcfour,des,md5,modes,sha1,sha2}}/sun4v/. and usr/src/common/bignum/sun4v/. Solaris 11 binary is available from the Oracle Solaris 11 download website. OpenSSL t4 engine The source for the OpenSSL t4 engine, which is based on the Solaris source above, is viewable through the OpenGrok source code browser in directory src/components/openssl/openssl-1.0.0/engines/t4 . You can download the source from the same website or through Mercurial source code management, hg(1). Conclusion Oracle Solaris with SPARC T4 provides a rich set of accelerated cryptographic and hash algorithms. Using the latest update, Solaris 11.1, provides the best set of optimized algorithms, but alternatives are often available, sometimes slightly slower, for releases back to Solaris 10 08/11 (U10). Reference See also these earlier blogs. SPARC T4 OpenSSL Engine by myself, Dan Anderson (2011), discusses the Openssl T4 engine and reviews the SPARC T4 processor for the Solaris 11 release. Exciting Crypto Advances with the T4 processor and Oracle Solaris 11 by Valerie Fenwick (2011) discusses crypto algorithms that were optimized for the T4 processor with the Solaris 11 FCS (11/11) and Solaris 10 08/11 (U10) release. T4 Crypto Cheat Sheet by Stefan Hinker (2012) discusses how to make T4 crypto optimization available to various consumers (such as SSH, Java, OpenSSL, Apache, etc.) High Performance Security For Oracle Database and Fusion Middleware Applications using SPARC T4 (PDF, 2012) discusses SPARC T4 and its usage to optimize application security. Configuring Oracle iPlanet WebServer / Oracle Traffic Director to use crypto accelerators on T4-1 servers by Meena Vyas (2012)

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  • Spooling in SQL execution plans

    - by Rob Farley
    Sewing has never been my thing. I barely even know the terminology, and when discussing this with American friends, I even found out that half the words that Americans use are different to the words that English and Australian people use. That said – let’s talk about spools! In particular, the Spool operators that you find in some SQL execution plans. This post is for T-SQL Tuesday, hosted this month by me! I’ve chosen to write about spools because they seem to get a bad rap (even in my song I used the line “There’s spooling from a CTE, they’ve got recursion needlessly”). I figured it was worth covering some of what spools are about, and hopefully explain why they are remarkably necessary, and generally very useful. If you have a look at the Books Online page about Plan Operators, at http://msdn.microsoft.com/en-us/library/ms191158.aspx, and do a search for the word ‘spool’, you’ll notice it says there are 46 matches. 46! Yeah, that’s what I thought too... Spooling is mentioned in several operators: Eager Spool, Lazy Spool, Index Spool (sometimes called a Nonclustered Index Spool), Row Count Spool, Spool, Table Spool, and Window Spool (oh, and Cache, which is a special kind of spool for a single row, but as it isn’t used in SQL 2012, I won’t describe it any further here). Spool, Table Spool, Index Spool, Window Spool and Row Count Spool are all physical operators, whereas Eager Spool and Lazy Spool are logical operators, describing the way that the other spools work. For example, you might see a Table Spool which is either Eager or Lazy. A Window Spool can actually act as both, as I’ll mention in a moment. In sewing, cotton is put onto a spool to make it more useful. You might buy it in bulk on a cone, but if you’re going to be using a sewing machine, then you quite probably want to have it on a spool or bobbin, which allows it to be used in a more effective way. This is the picture that I want you to think about in relation to your data. I’m sure you use spools every time you use your sewing machine. I know I do. I can’t think of a time when I’ve got out my sewing machine to do some sewing and haven’t used a spool. However, I often run SQL queries that don’t use spools. You see, the data that is consumed by my query is typically in a useful state without a spool. It’s like I can just sew with my cotton despite it not being on a spool! Many of my favourite features in T-SQL do like to use spools though. This looks like a very similar query to before, but includes an OVER clause to return a column telling me the number of rows in my data set. I’ll describe what’s going on in a few paragraphs’ time. So what does a Spool operator actually do? The spool operator consumes a set of data, and stores it in a temporary structure, in the tempdb database. This structure is typically either a Table (ie, a heap), or an Index (ie, a b-tree). If no data is actually needed from it, then it could also be a Row Count spool, which only stores the number of rows that the spool operator consumes. A Window Spool is another option if the data being consumed is tightly linked to windows of data, such as when the ROWS/RANGE clause of the OVER clause is being used. You could maybe think about the type of spool being like whether the cotton is going onto a small bobbin to fit in the base of the sewing machine, or whether it’s a larger spool for the top. A Table or Index Spool is either Eager or Lazy in nature. Eager and Lazy are Logical operators, which talk more about the behaviour, rather than the physical operation. If I’m sewing, I can either be all enthusiastic and get all my cotton onto the spool before I start, or I can do it as I need it. “Lazy” might not the be the best word to describe a person – in the SQL world it describes the idea of either fetching all the rows to build up the whole spool when the operator is called (Eager), or populating the spool only as it’s needed (Lazy). Window Spools are both physical and logical. They’re eager on a per-window basis, but lazy between windows. And when is it needed? The way I see it, spools are needed for two reasons. 1 – When data is going to be needed AGAIN. 2 – When data needs to be kept away from the original source. If you’re someone that writes long stored procedures, you are probably quite aware of the second scenario. I see plenty of stored procedures being written this way – where the query writer populates a temporary table, so that they can make updates to it without risking the original table. SQL does this too. Imagine I’m updating my contact list, and some of my changes move data to later in the book. If I’m not careful, I might update the same row a second time (or even enter an infinite loop, updating it over and over). A spool can make sure that I don’t, by using a copy of the data. This problem is known as the Halloween Effect (not because it’s spooky, but because it was discovered in late October one year). As I’m sure you can imagine, the kind of spool you’d need to protect against the Halloween Effect would be eager, because if you’re only handling one row at a time, then you’re not providing the protection... An eager spool will block the flow of data, waiting until it has fetched all the data before serving it up to the operator that called it. In the query below I’m forcing the Query Optimizer to use an index which would be upset if the Name column values got changed, and we see that before any data is fetched, a spool is created to load the data into. This doesn’t stop the index being maintained, but it does mean that the index is protected from the changes that are being done. There are plenty of times, though, when you need data repeatedly. Consider the query I put above. A simple join, but then counting the number of rows that came through. The way that this has executed (be it ideal or not), is to ask that a Table Spool be populated. That’s the Table Spool operator on the top row. That spool can produce the same set of rows repeatedly. This is the behaviour that we see in the bottom half of the plan. In the bottom half of the plan, we see that the a join is being done between the rows that are being sourced from the spool – one being aggregated and one not – producing the columns that we need for the query. Table v Index When considering whether to use a Table Spool or an Index Spool, the question that the Query Optimizer needs to answer is whether there is sufficient benefit to storing the data in a b-tree. The idea of having data in indexes is great, but of course there is a cost to maintaining them. Here we’re creating a temporary structure for data, and there is a cost associated with populating each row into its correct position according to a b-tree, as opposed to simply adding it to the end of the list of rows in a heap. Using a b-tree could even result in page-splits as the b-tree is populated, so there had better be a reason to use that kind of structure. That all depends on how the data is going to be used in other parts of the plan. If you’ve ever thought that you could use a temporary index for a particular query, well this is it – and the Query Optimizer can do that if it thinks it’s worthwhile. It’s worth noting that just because a Spool is populated using an Index Spool, it can still be fetched using a Table Spool. The details about whether or not a Spool used as a source shows as a Table Spool or an Index Spool is more about whether a Seek predicate is used, rather than on the underlying structure. Recursive CTE I’ve already shown you an example of spooling when the OVER clause is used. You might see them being used whenever you have data that is needed multiple times, and CTEs are quite common here. With the definition of a set of data described in a CTE, if the query writer is leveraging this by referring to the CTE multiple times, and there’s no simplification to be leveraged, a spool could theoretically be used to avoid reapplying the CTE’s logic. Annoyingly, this doesn’t happen. Consider this query, which really looks like it’s using the same data twice. I’m creating a set of data (which is completely deterministic, by the way), and then joining it back to itself. There seems to be no reason why it shouldn’t use a spool for the set described by the CTE, but it doesn’t. On the other hand, if we don’t pull as many columns back, we might see a very different plan. You see, CTEs, like all sub-queries, are simplified out to figure out the best way of executing the whole query. My example is somewhat contrived, and although there are plenty of cases when it’s nice to give the Query Optimizer hints about how to execute queries, it usually doesn’t do a bad job, even without spooling (and you can always use a temporary table). When recursion is used, though, spooling should be expected. Consider what we’re asking for in a recursive CTE. We’re telling the system to construct a set of data using an initial query, and then use set as a source for another query, piping this back into the same set and back around. It’s very much a spool. The analogy of cotton is long gone here, as the idea of having a continual loop of cotton feeding onto a spool and off again doesn’t quite fit, but that’s what we have here. Data is being fed onto the spool, and getting pulled out a second time when the spool is used as a source. (This query is running on AdventureWorks, which has a ManagerID column in HumanResources.Employee, not AdventureWorks2012) The Index Spool operator is sucking rows into it – lazily. It has to be lazy, because at the start, there’s only one row to be had. However, as rows get populated onto the spool, the Table Spool operator on the right can return rows when asked, ending up with more rows (potentially) getting back onto the spool, ready for the next round. (The Assert operator is merely checking to see if we’ve reached the MAXRECURSION point – it vanishes if you use OPTION (MAXRECURSION 0), which you can try yourself if you like). Spools are useful. Don’t lose sight of that. Every time you use temporary tables or table variables in a stored procedure, you’re essentially doing the same – don’t get upset at the Query Optimizer for doing so, even if you think the spool looks like an expensive part of the query. I hope you’re enjoying this T-SQL Tuesday. Why not head over to my post that is hosting it this month to read about some other plan operators? At some point I’ll write a summary post – once I have you should find a comment below pointing at it. @rob_farley

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  • Constant game speed independent of variable FPS in OpenGL with GLUT?

    - by Nazgulled
    I've been reading Koen Witters detailed article about different game loop solutions but I'm having some problems implementing the last one with GLUT, which is the recommended one. After reading a couple of articles, tutorials and code from other people on how to achieve a constant game speed, I think that what I currently have implemented (I'll post the code below) is what Koen Witters called Game Speed dependent on Variable FPS, the second on his article. First, through my searching experience, there's a couple of people that probably have the knowledge to help out on this but don't know what GLUT is and I'm going to try and explain (feel free to correct me) the relevant functions for my problem of this OpenGL toolkit. Skip this section if you know what GLUT is and how to play with it. GLUT Toolkit: GLUT is an OpenGL toolkit and helps with common tasks in OpenGL. The glutDisplayFunc(renderScene) takes a pointer to a renderScene() function callback, which will be responsible for rendering everything. The renderScene() function will only be called once after the callback registration. The glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0) takes the number of milliseconds to pass before calling the callback processAnimationTimer(). The last argument is just a value to pass to the timer callback. The processAnimationTimer() will not be called each TIMER_MILLISECONDS but just once. The glutPostRedisplay() function requests GLUT to render a new frame so we need call this every time we change something in the scene. The glutIdleFunc(renderScene) could be used to register a callback to renderScene() (this does not make glutDisplayFunc() irrelevant) but this function should be avoided because the idle callback is continuously called when events are not being received, increasing the CPU load. The glutGet(GLUT_ELAPSED_TIME) function returns the number of milliseconds since glutInit was called (or first call to glutGet(GLUT_ELAPSED_TIME)). That's the timer we have with GLUT. I know there are better alternatives for high resolution timers, but let's keep with this one for now. I think this is enough information on how GLUT renders frames so people that didn't know about it could also pitch in this question to try and help if they fell like it. Current Implementation: Now, I'm not sure I have correctly implemented the second solution proposed by Koen, Game Speed dependent on Variable FPS. The relevant code for that goes like this: #define TICKS_PER_SECOND 30 #define MOVEMENT_SPEED 2.0f const int TIMER_MILLISECONDS = 1000 / TICKS_PER_SECOND; int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void processAnimationTimer(int value) { // setups the timer to be called again glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Requests to render a new frame (this will call my renderScene() once) glutPostRedisplay(); } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) // Setup the timer to be called one first time glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Read the current time since glutInit was called currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } This implementation doesn't fell right. It works in the sense that helps the game speed to be constant dependent on the FPS. So that moving from point A to point B takes the same time no matter the high/low framerate. However, I believe I'm limiting the game framerate with this approach. Each frame will only be rendered when the time callback is called, that means the framerate will be roughly around TICKS_PER_SECOND frames per second. This doesn't feel right, you shouldn't limit your powerful hardware, it's wrong. It's my understanding though, that I still need to calculate the elapsedTime. Just because I'm telling GLUT to call the timer callback every TIMER_MILLISECONDS, it doesn't mean it will always do that on time. I'm not sure how can I fix this and to be completely honest, I have no idea what is the game loop in GLUT, you know, the while( game_is_running ) loop in Koen's article. But it's my understanding that GLUT is event-driven and that game loop starts when I call glutMainLoop() (which never returns), yes? I thought I could register an idle callback with glutIdleFunc() and use that as replacement of glutTimerFunc(), only rendering when necessary (instead of all the time as usual) but when I tested this with an empty callback (like void gameLoop() {}) and it was basically doing nothing, only a black screen, the CPU spiked to 25% and remained there until I killed the game and it went back to normal. So I don't think that's the path to follow. Using glutTimerFunc() is definitely not a good approach to perform all movements/animations based on that, as I'm limiting my game to a constant FPS, not cool. Or maybe I'm using it wrong and my implementation is not right? How exactly can I have a constant game speed with variable FPS? More exactly, how do I correctly implement Koen's Constant Game Speed with Maximum FPS solution (the fourth one on his article) with GLUT? Maybe this is not possible at all with GLUT? If not, what are my alternatives? What is the best approach to this problem (constant game speed) with GLUT? I originally posted this question on Stack Overflow before being pointed out about this site. The following is a different approach I tried after creating the question in SO, so I'm posting it here too. Another Approach: I've been experimenting and here's what I was able to achieve now. Instead of calculating the elapsed time on a timed function (which limits my game's framerate) I'm now doing it in renderScene(). Whenever changes to the scene happen I call glutPostRedisplay() (ie: camera moving, some object animation, etc...) which will make a call to renderScene(). I can use the elapsed time in this function to move my camera for instance. My code has now turned into this: int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void renderScene(void) { (...) // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Setup the camera position and looking point SceneCamera.LookAt(); // All drawing code goes inside this function drawCompleteScene(); glutSwapBuffers(); /* Redraw the frame ONLY if the user is moving the camera (similar code will be needed to redraw the frame for other events) */ if(!IsTupleEmpty(cameraDirection)) { glutPostRedisplay(); } } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } Conclusion, it's working, or so it seems. If I don't move the camera, the CPU usage is low, nothing is being rendered (for testing purposes I only have a grid extending for 4000.0f, while zFar is set to 1000.0f). When I start moving the camera the scene starts redrawing itself. If I keep pressing the move keys, the CPU usage will increase; this is normal behavior. It drops back when I stop moving. Unless I'm missing something, it seems like a good approach for now. I did find this interesting article on iDevGames and this implementation is probably affected by the problem described on that article. What's your thoughts on that? Please note that I'm just doing this for fun, I have no intentions of creating some game to distribute or something like that, not in the near future at least. If I did, I would probably go with something else besides GLUT. But since I'm using GLUT, and other than the problem described on iDevGames, do you think this latest implementation is sufficient for GLUT? The only real issue I can think of right now is that I'll need to keep calling glutPostRedisplay() every time the scene changes something and keep calling it until there's nothing new to redraw. A little complexity added to the code for a better cause, I think. What do you think?

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  • Taming Hopping Windows

    - by Roman Schindlauer
    At first glance, hopping windows seem fairly innocuous and obvious. They organize events into windows with a simple periodic definition: the windows have some duration d (e.g. a window covers 5 second time intervals), an interval or period p (e.g. a new window starts every 2 seconds) and an alignment a (e.g. one of those windows starts at 12:00 PM on March 15, 2012 UTC). var wins = xs     .HoppingWindow(TimeSpan.FromSeconds(5),                    TimeSpan.FromSeconds(2),                    new DateTime(2012, 3, 15, 12, 0, 0, DateTimeKind.Utc)); Logically, there is a window with start time a + np and end time a + np + d for every integer n. That’s a lot of windows. So why doesn’t the following query (always) blow up? var query = wins.Select(win => win.Count()); A few users have asked why StreamInsight doesn’t produce output for empty windows. Primarily it’s because there is an infinite number of empty windows! (Actually, StreamInsight uses DateTimeOffset.MaxValue to approximate “the end of time” and DateTimeOffset.MinValue to approximate “the beginning of time”, so the number of windows is lower in practice.) That was the good news. Now the bad news. Events also have duration. Consider the following simple input: var xs = this.Application                 .DefineEnumerable(() => new[]                     { EdgeEvent.CreateStart(DateTimeOffset.UtcNow, 0) })                 .ToStreamable(AdvanceTimeSettings.IncreasingStartTime); Because the event has no explicit end edge, it lasts until the end of time. So there are lots of non-empty windows if we apply a hopping window to that single event! For this reason, we need to be careful with hopping window queries in StreamInsight. Or we can switch to a custom implementation of hopping windows that doesn’t suffer from this shortcoming. The alternate window implementation produces output only when the input changes. We start by breaking up the timeline into non-overlapping intervals assigned to each window. In figure 1, six hopping windows (“Windows”) are assigned to six intervals (“Assignments”) in the timeline. Next we take input events (“Events”) and alter their lifetimes (“Altered Events”) so that they cover the intervals of the windows they intersect. In figure 1, you can see that the first event e1 intersects windows w1 and w2 so it is adjusted to cover assignments a1 and a2. Finally, we can use snapshot windows (“Snapshots”) to produce output for the hopping windows. Notice however that instead of having six windows generating output, we have only four. The first and second snapshots correspond to the first and second hopping windows. The remaining snapshots however cover two hopping windows each! While in this example we saved only two events, the savings can be more significant when the ratio of event duration to window duration is higher. Figure 1: Timeline The implementation of this strategy is straightforward. We need to set the start times of events to the start time of the interval assigned to the earliest window including the start time. Similarly, we need to modify the end times of events to the end time of the interval assigned to the latest window including the end time. The following snap-to-boundary function that rounds a timestamp value t down to the nearest value t' <= t such that t' is a + np for some integer n will be useful. For convenience, we will represent both DateTime and TimeSpan values using long ticks: static long SnapToBoundary(long t, long a, long p) {     return t - ((t - a) % p) - (t > a ? 0L : p); } How do we find the earliest window including the start time for an event? It’s the window following the last window that does not include the start time assuming that there are no gaps in the windows (i.e. duration < interval), and limitation of this solution. To find the end time of that antecedent window, we need to know the alignment of window ends: long e = a + (d % p); Using the window end alignment, we are finally ready to describe the start time selector: static long AdjustStartTime(long t, long e, long p) {     return SnapToBoundary(t, e, p) + p; } To find the latest window including the end time for an event, we look for the last window start time (non-inclusive): public static long AdjustEndTime(long t, long a, long d, long p) {     return SnapToBoundary(t - 1, a, p) + p + d; } Bringing it together, we can define the translation from events to ‘altered events’ as in Figure 1: public static IQStreamable<T> SnapToWindowIntervals<T>(IQStreamable<T> source, TimeSpan duration, TimeSpan interval, DateTime alignment) {     if (source == null) throw new ArgumentNullException("source");     // reason about DateTime and TimeSpan in ticks     long d = Math.Min(DateTime.MaxValue.Ticks, duration.Ticks);     long p = Math.Min(DateTime.MaxValue.Ticks, Math.Abs(interval.Ticks));     // set alignment to earliest possible window     var a = alignment.ToUniversalTime().Ticks % p;     // verify constraints of this solution     if (d <= 0L) { throw new ArgumentOutOfRangeException("duration"); }     if (p == 0L || p > d) { throw new ArgumentOutOfRangeException("interval"); }     // find the alignment of window ends     long e = a + (d % p);     return source.AlterEventLifetime(         evt => ToDateTime(AdjustStartTime(evt.StartTime.ToUniversalTime().Ticks, e, p)),         evt => ToDateTime(AdjustEndTime(evt.EndTime.ToUniversalTime().Ticks, a, d, p)) -             ToDateTime(AdjustStartTime(evt.StartTime.ToUniversalTime().Ticks, e, p))); } public static DateTime ToDateTime(long ticks) {     // just snap to min or max value rather than under/overflowing     return ticks < DateTime.MinValue.Ticks         ? new DateTime(DateTime.MinValue.Ticks, DateTimeKind.Utc)         : ticks > DateTime.MaxValue.Ticks         ? new DateTime(DateTime.MaxValue.Ticks, DateTimeKind.Utc)         : new DateTime(ticks, DateTimeKind.Utc); } Finally, we can describe our custom hopping window operator: public static IQWindowedStreamable<T> HoppingWindow2<T>(     IQStreamable<T> source,     TimeSpan duration,     TimeSpan interval,     DateTime alignment) {     if (source == null) { throw new ArgumentNullException("source"); }     return SnapToWindowIntervals(source, duration, interval, alignment).SnapshotWindow(); } By switching from HoppingWindow to HoppingWindow2 in the following example, the query returns quickly rather than gobbling resources and ultimately failing! public void Main() {     var start = new DateTimeOffset(new DateTime(2012, 6, 28), TimeSpan.Zero);     var duration = TimeSpan.FromSeconds(5);     var interval = TimeSpan.FromSeconds(2);     var alignment = new DateTime(2012, 3, 15, 12, 0, 0, DateTimeKind.Utc);     var events = this.Application.DefineEnumerable(() => new[]     {         EdgeEvent.CreateStart(start.AddSeconds(0), "e0"),         EdgeEvent.CreateStart(start.AddSeconds(1), "e1"),         EdgeEvent.CreateEnd(start.AddSeconds(1), start.AddSeconds(2), "e1"),         EdgeEvent.CreateStart(start.AddSeconds(3), "e2"),         EdgeEvent.CreateStart(start.AddSeconds(9), "e3"),         EdgeEvent.CreateEnd(start.AddSeconds(3), start.AddSeconds(10), "e2"),         EdgeEvent.CreateEnd(start.AddSeconds(9), start.AddSeconds(10), "e3"),     }).ToStreamable(AdvanceTimeSettings.IncreasingStartTime);     var adjustedEvents = SnapToWindowIntervals(events, duration, interval, alignment);     var query = from win in HoppingWindow2(events, duration, interval, alignment)                 select win.Count();     DisplayResults(adjustedEvents, "Adjusted Events");     DisplayResults(query, "Query"); } As you can see, instead of producing a massive number of windows for the open start edge e0, a single window is emitted from 12:00:15 AM until the end of time: Adjusted Events StartTime EndTime Payload 6/28/2012 12:00:01 AM 12/31/9999 11:59:59 PM e0 6/28/2012 12:00:03 AM 6/28/2012 12:00:07 AM e1 6/28/2012 12:00:05 AM 6/28/2012 12:00:15 AM e2 6/28/2012 12:00:11 AM 6/28/2012 12:00:15 AM e3 Query StartTime EndTime Payload 6/28/2012 12:00:01 AM 6/28/2012 12:00:03 AM 1 6/28/2012 12:00:03 AM 6/28/2012 12:00:05 AM 2 6/28/2012 12:00:05 AM 6/28/2012 12:00:07 AM 3 6/28/2012 12:00:07 AM 6/28/2012 12:00:11 AM 2 6/28/2012 12:00:11 AM 6/28/2012 12:00:15 AM 3 6/28/2012 12:00:15 AM 12/31/9999 11:59:59 PM 1 Regards, The StreamInsight Team

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  • Combine 3D objects in XNA 4

    - by Christoph
    Currently I am writing on my thesis for university, the theme I am working on is 3D Visualization of hierarchical structures using cone trees. I want to do is to draw a cone and arrange a number of spheres at the bottom of the cone. The spheres should be arranged according to the radius and the number of spheres correctly. As you can imagine I need a lot of these cone/sphere combinations. First Attempt I was able to find some tutorials that helped with drawing cones and spheres. Cone public Cone(GraphicsDevice device, float height, int tessellation, string name, List<Sphere> children) { //prepare children and calculate the children spacing and radius of the cone if (children == null || children.Count == 0) { throw new ArgumentNullException("children"); } this.Height = height; this.Name = name; this.Children = children; //create the cone if (tessellation < 3) { throw new ArgumentOutOfRangeException("tessellation"); } //Create a ring of triangels around the outside of the cones bottom for (int i = 0; i < tessellation; i++) { Vector3 normal = this.GetCircleVector(i, tessellation); // add the vertices for the top of the cone base.AddVertex(Vector3.Up * height, normal); //add the bottom circle base.AddVertex(normal * this.radius + Vector3.Down * height, normal); //Add indices base.AddIndex(i * 2); base.AddIndex(i * 2 + 1); base.AddIndex((i * 2 + 2) % (tessellation * 2)); base.AddIndex(i * 2 + 1); base.AddIndex((i * 2 + 3) % (tessellation * 2)); base.AddIndex((i * 2 + 2) % (tessellation * 2)); } //create flate triangle to seal the bottom this.CreateCap(tessellation, height, this.Radius, Vector3.Down); base.InitializePrimitive(device); } Sphere public void Initialize(GraphicsDevice device, Vector3 qi) { int verticalSegments = this.Tesselation; int horizontalSegments = this.Tesselation * 2; //single vertex on the bottom base.AddVertex((qi * this.Radius) + this.lowering, Vector3.Down); for (int i = 0; i < verticalSegments; i++) { float latitude = ((i + 1) * MathHelper.Pi / verticalSegments) - MathHelper.PiOver2; float dy = (float)Math.Sin(latitude); float dxz = (float)Math.Cos(latitude); //Create a singe ring of latitudes for (int j = 0; j < horizontalSegments; j++) { float longitude = j * MathHelper.TwoPi / horizontalSegments; float dx = (float)Math.Cos(longitude) * dxz; float dz = (float)Math.Sin(longitude) * dxz; Vector3 normal = new Vector3(dx, dy, dz); base.AddVertex(normal * this.Radius, normal); } } // Finish with a single vertex at the top of the sphere. AddVertex((qi * this.Radius) + this.lowering, Vector3.Up); // Create a fan connecting the bottom vertex to the bottom latitude ring. for (int i = 0; i < horizontalSegments; i++) { AddIndex(0); AddIndex(1 + (i + 1) % horizontalSegments); AddIndex(1 + i); } // Fill the sphere body with triangles joining each pair of latitude rings. for (int i = 0; i < verticalSegments - 2; i++) { for (int j = 0; j < horizontalSegments; j++) { int nextI = i + 1; int nextJ = (j + 1) % horizontalSegments; base.AddIndex(1 + i * horizontalSegments + j); base.AddIndex(1 + i * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + j); base.AddIndex(1 + i * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + j); } } // Create a fan connecting the top vertex to the top latitude ring. for (int i = 0; i < horizontalSegments; i++) { base.AddIndex(CurrentVertex - 1); base.AddIndex(CurrentVertex - 2 - (i + 1) % horizontalSegments); base.AddIndex(CurrentVertex - 2 - i); } base.InitializePrimitive(device); } The tricky part now is to arrange the spheres at the bottom of the cone. I tried is to draw just the cone and then draw the spheres. I need a lot of these cones, so it would be pretty hard to calculate all the positions correctly. Second Attempt So the second try was to generate a object that builds all vertices of the cone and all of the spheres at once. So I was hoping to render a cone with all its spheres arranged correctly. After a short debug I found out that the cone is created and the first sphere, when it turn of the second sphere I am running into an OutOfBoundsException of ushort.MaxValue. Cone and Spheres public ConeWithSpheres(GraphicsDevice device, float height, float coneDiameter, float sphereDiameter, int coneTessellation, int sphereTessellation, int numberOfSpheres) { if (coneTessellation < 3) { throw new ArgumentException(string.Format("{0} is to small for the tessellation of the cone. The number must be greater or equal to 3", coneTessellation)); } if (sphereTessellation < 3) { throw new ArgumentException(string.Format("{0} is to small for the tessellation of the sphere. The number must be greater or equal to 3", sphereTessellation)); } //set properties this.Height = height; this.ConeDiameter = coneDiameter; this.SphereDiameter = sphereDiameter; this.NumberOfChildren = numberOfSpheres; //end set properties //generate the cone this.GenerateCone(device, coneTessellation); //generate the spheres //vector that defines the Y position of the sphere on the cones bottom Vector3 lowering = new Vector3(0, 0.888f, 0); this.GenerateSpheres(device, sphereTessellation, numberOfSpheres, lowering); } // ------ GENERATE CONE ------ private void GenerateCone(GraphicsDevice device, int coneTessellation) { int doubleTessellation = coneTessellation * 2; //Create a ring of triangels around the outside of the cones bottom for (int index = 0; index < coneTessellation; index++) { Vector3 normal = this.GetCircleVector(index, coneTessellation); //add the vertices for the top of the cone base.AddVertex(Vector3.Up * this.Height, normal); //add the bottom of the cone base.AddVertex(normal * this.ConeRadius + Vector3.Down * this.Height, normal); //add indices base.AddIndex(index * 2); base.AddIndex(index * 2 + 1); base.AddIndex((index * 2 + 2) % doubleTessellation); base.AddIndex(index * 2 + 1); base.AddIndex((index * 2 + 3) % doubleTessellation); base.AddIndex((index * 2 + 2) % doubleTessellation); } //create flate triangle to seal the bottom this.CreateCap(coneTessellation, this.Height, this.ConeRadius, Vector3.Down); base.InitializePrimitive(device); } // ------ GENERATE SPHERES ------ private void GenerateSpheres(GraphicsDevice device, int sphereTessellation, int numberOfSpheres, Vector3 lowering) { int verticalSegments = sphereTessellation; int horizontalSegments = sphereTessellation * 2; for (int childCount = 1; childCount < numberOfSpheres; childCount++) { //single vertex at the bottom of the sphere base.AddVertex((this.GetCircleVector(childCount, this.NumberOfChildren) * this.SphereRadius) + lowering, Vector3.Down); for (int verticalSegmentsCount = 0; verticalSegmentsCount < verticalSegments; verticalSegmentsCount++) { float latitude = ((verticalSegmentsCount + 1) * MathHelper.Pi / verticalSegments) - MathHelper.PiOver2; float dy = (float)Math.Sin(latitude); float dxz = (float)Math.Cos(latitude); //create a single ring of latitudes for (int horizontalSegmentsCount = 0; horizontalSegmentsCount < horizontalSegments; horizontalSegmentsCount++) { float longitude = horizontalSegmentsCount * MathHelper.TwoPi / horizontalSegments; float dx = (float)Math.Cos(longitude) * dxz; float dz = (float)Math.Sin(longitude) * dxz; Vector3 normal = new Vector3(dx, dy, dz); base.AddVertex((normal * this.SphereRadius) + lowering, normal); } } //finish with a single vertex at the top of the sphere base.AddVertex((this.GetCircleVector(childCount, this.NumberOfChildren) * this.SphereRadius) + lowering, Vector3.Up); //create a fan connecting the bottom vertex to the bottom latitude ring for (int i = 0; i < horizontalSegments; i++) { base.AddIndex(0); base.AddIndex(1 + (i + 1) % horizontalSegments); base.AddIndex(1 + i); } //Fill the sphere body with triangles joining each pair of latitude rings for (int i = 0; i < verticalSegments - 2; i++) { for (int j = 0; j < horizontalSegments; j++) { int nextI = i + 1; int nextJ = (j + 1) % horizontalSegments; base.AddIndex(1 + i * horizontalSegments + j); base.AddIndex(1 + i * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + j); base.AddIndex(1 + i * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + nextJ); base.AddIndex(1 + nextI * horizontalSegments + j); } } //create a fan connecting the top vertiex to the top latitude for (int i = 0; i < horizontalSegments; i++) { base.AddIndex(this.CurrentVertex - 1); base.AddIndex(this.CurrentVertex - 2 - (i + 1) % horizontalSegments); base.AddIndex(this.CurrentVertex - 2 - i); } base.InitializePrimitive(device); } } Any ideas how I could fix this?

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  • Adventures in Windows 8: Placing items in a GridView with a ColumnSpan or RowSpan

    - by Laurent Bugnion
    Currently working on a Windows 8 app for an important client, I will be writing about small issues, tips and tricks, ideas and whatever occurs to me during the development and the integration of this app. When working with a GridView, it is quite common to use a VariableSizedWrapGrid as the ItemsPanel. This creates a nice flowing layout which will auto-adapt for various resolutions. This is ideal when you want to build views like the Windows 8 start menu. However immediately we notice that the Start menu allows to place items on one column (Smaller) or two columns (Larger). This switch happens through the AppBar. So how do we implement that in our app? Using ColumnSpan and RowSpan When you use a VariableSizedWrapGrid directly in your XAML, you can attach the VariableSizedWrapGrid.ColumnSpan and VariableSizedWrapGrid.RowSpan attached properties directly to an item to create the desired effect. For instance this code create this output (shown in Blend but it runs just the same): <VariableSizedWrapGrid ItemHeight="100" ItemWidth="100" Width="200" Orientation="Horizontal"> <Rectangle Fill="Purple" /> <Rectangle Fill="Orange" /> <Rectangle Fill="Yellow" VariableSizedWrapGrid.ColumnSpan="2" /> <Rectangle Fill="Red" VariableSizedWrapGrid.ColumnSpan="2" VariableSizedWrapGrid.RowSpan="2" /> <Rectangle Fill="Green" VariableSizedWrapGrid.RowSpan="2" /> <Rectangle Fill="Blue" /> <Rectangle Fill="LightGray" /> </VariableSizedWrapGrid> Using the VariableSizedWrapGrid as ItemsPanel When you use a GridView however, you typically bind the ItemsSource property to a collection, for example in a viewmodel. In that case, you want to be able to switch the ColumnSpan and RowSpan depending on properties on the item. I tried to find a way to bind the VariableSizedWrapGrid.ColumnSpan attached property on the GridView’s ItemContainerStyle template to an observable property on the item, but it didn’t work. Instead, I decided to use a StyleSelector to switch the GridViewItem’s style. Here’s how: First I added my two GridViews to my XAML as follows: <Page.Resources> <local:MainViewModel x:Key="Main" /> <DataTemplate x:Key="DataTemplate1"> <Grid Background="{Binding Brush}"> <TextBlock Text="{Binding BrushCode}" /> </Grid> </DataTemplate> </Page.Resources> <Page.DataContext> <Binding Source="{StaticResource Main}" /> </Page.DataContext> <Grid Background="{StaticResource ApplicationPageBackgroundThemeBrush}" Margin="20"> <Grid.ColumnDefinitions> <ColumnDefinition Width="Auto" /> <ColumnDefinition Width="*" /> </Grid.ColumnDefinitions> <GridView ItemsSource="{Binding Items}" ItemTemplate="{StaticResource DataTemplate1}" VerticalAlignment="Top"> <GridView.ItemsPanel> <ItemsPanelTemplate> <VariableSizedWrapGrid ItemHeight="150" ItemWidth="150" /> </ItemsPanelTemplate> </GridView.ItemsPanel> </GridView> <GridView Grid.Column="1" ItemsSource="{Binding Items}" ItemTemplate="{StaticResource DataTemplate1}" VerticalAlignment="Top"> <GridView.ItemsPanel> <ItemsPanelTemplate> <VariableSizedWrapGrid ItemHeight="100" ItemWidth="100" /> </ItemsPanelTemplate> </GridView.ItemsPanel> </GridView> </Grid> The MainViewModel looks like this: public class MainViewModel { public IList<Item> Items { get; private set; } public MainViewModel() { Items = new List<Item> { new Item { Brush = new SolidColorBrush(Colors.Red) }, new Item { Brush = new SolidColorBrush(Colors.Blue) }, new Item { Brush = new SolidColorBrush(Colors.Green), }, // And more... }; } } As for the Item class, I am using an MVVM Light ObservableObject but you can use your own simple implementation of INotifyPropertyChanged of course: public class Item : ObservableObject { public const string ColSpanPropertyName = "ColSpan"; private int _colSpan = 1; public int ColSpan { get { return _colSpan; } set { Set(ColSpanPropertyName, ref _colSpan, value); } } public SolidColorBrush Brush { get; set; } public string BrushCode { get { return Brush.Color.ToString(); } } } Then I copied the GridViewItem’s style locally. To do this, I use Expression Blend’s functionality. It has the disadvantage to copy a large portion of XAML to your application, but the HUGE advantage to allow you to change the look and feel of your GridViewItem everywhere in the application. For example, you can change the selection chrome, the item’s alignments and many other properties. Actually everytime I use a ListBox, ListView or any other data control, I typically copy the item style to a resource dictionary in my application and I tweak it. Note that Blend for Windows 8 apps is automatically installed with every edition of Visual Studio 2012 (including Express) so you have no excuses anymore not to use Blend :) Open MainPage.xaml in Expression Blend by right clicking on the MainPage.xaml file in the Solution Explorer and selecting Open in Blend from the context menu. Note that the items do not look very nice! The reason is that the default ItemContainerStyle sets the content’s alignment to “Center” which I never quite understood. Seems to me that you rather want the content to be stretched, but anyway it is easy to change.   Right click on the GridView on the left and select Edit Additional Templates, Edit Generated Item Container (ItemContainerStyle), Edit a Copy. In the Create Style Resource dialog, enter the name “DefaultGridViewItemStyle”, select “Application” and press OK. Side note 1: You need to save in a global resource dictionary because later we will need to retrieve that Style from a global location. Side note 2": I would rather copy the style to an external resource dictionary that I link into the App.xaml file, but I want to keep things simple here. Blend switches in Template edit mode. The template you are editing now is inside the ItemContainerStyle and will govern the appearance of your items. This is where, for instance, the “checked” chrome is defined, and where you can alter it if you need to. Note that you can reuse this style for all your GridViews even if you use a different DataTemplate for your items. Makes sense? I probably need to think about writing another blog post dedicated to the ItemContainerStyle :) In the breadcrumb bar on top of the page, click on the style icon. The property we want to change now can be changed in the Style instead of the Template, which is a better idea. Blend is not in Style edit mode, as you can see in the Objects and Timeline pane. In the Properties pane, in the Search box, enter the word “content”. This will filter all the properties containing that partial string, including the two we are interested in: HorizontalContentAlignment and VerticalContentAlignment. Set these two values to “Stretch” instead of the default “Center”. Using the breadcrumb bar again, set the scope back to the Page (by clicking on the first crumb on the left). Notice how the items are now showing as squares in the first GridView. We will now use the same ItemContainerStyle for the second GridView. To do this, right click on the second GridView and select Edit Additional Templates, Edit Generate Item Container, Apply Resource, DefaultGridViewItemStyle. The page now looks nicer: And now for the ColumnSpan! So now, let’s change the ColumnSpan property. First, let’s define a new Style that inherits the ItemContainerStyle we created before. Make sure that you save everything in Blend by pressing Ctrl-Shift-S. Open App.xaml in Visual Studio. Below the newly created DefaultGridViewItemStyle resource, add the following style: <Style x:Key="WideGridViewItemStyle" TargetType="GridViewItem" BasedOn="{StaticResource DefaultGridViewItemStyle}"> <Setter Property="VariableSizedWrapGrid.ColumnSpan" Value="2" /> </Style> Add a new class to the project, and name it MainItemStyleSelector. Implement the class as follows: public class MainItemStyleSelector : StyleSelector { protected override Style SelectStyleCore(object item, DependencyObject container) { var i = (Item)item; if (i.ColSpan == 2) { return Application.Current.Resources["WideGridViewItemStyle"] as Style; } return Application.Current.Resources["DefaultGridViewItemStyle"] as Style; } } In MainPage.xaml, add a resource to the Page.Resources section: <local:MainItemStyleSelector x:Key="MainItemStyleSelector" /> In MainPage.xaml, replace the ItemContainerStyle property on the first GridView with the ItemContainerStyleSelector property, pointing to the StaticResource we just defined. <GridView ItemsSource="{Binding Items}" ItemTemplate="{StaticResource DataTemplate1}" VerticalAlignment="Top" ItemContainerStyleSelector="{StaticResource MainItemStyleSelector}"> <GridView.ItemsPanel> <ItemsPanelTemplate> <VariableSizedWrapGrid ItemHeight="150" ItemWidth="150" /> </ItemsPanelTemplate> </GridView.ItemsPanel> </GridView> Do the same for the second GridView as well. Finally, in the MainViewModel, change the ColumnSpan property on the 3rd Item to 2. new Item { Brush = new SolidColorBrush(Colors.Green), ColSpan = 2 }, Running the application now creates the following image, which is what we wanted. Notice how the green item is now a “wide tile”. You can also experiment by creating different Styles, all inheriting the DefaultGridViewItemStyle and using different values of RowSpan for instance. This will allow you to create any layout you want, while leaving the heavy lifting of “flowing the layout” to the GridView control. What about changing these values dynamically? Of course as we can see in the Start menu, it would be nice to be able to change the ColumnSpan and maybe even the RowSpan values at runtime. Unfortunately at this time I have not found a good way to do that. I am investigating however and will make sure to post a follow up when I find what I am looking for!   Laurent Bugnion (GalaSoft) Subscribe | Twitter | Facebook | Flickr | LinkedIn

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  • Windows 7 cannot join samba domain

    - by Antonis Christofides
    I have a 3.5.6 samba server with a LDAP backend (both on Debian 6.0). I've been successfully adding Windows XP machines to the domain for years. I now try to add Windows 7. I have made the recommended registry changes, but I don't have any success so far. Here is what happens: 1. I go to computer name, select "Domain" instead of "Workgroup", type in the domain name, click OK. It asks me for the username and password of an account that can add computers to the domain; I enter them. After about 40 seconds, I get the following message: The following error occurred attempting to join the domain "ITIA": The specified computer account could not be found. Contact an administrator to verify the account is in the domain. If the account has been deleted unjoin, reboot, and rejoin the domain. Despite this, the samba server successfully creates the computer account. 2. Therefore, if I try again a second time, without deleting the already created computer account, I get a different error: The following error occurred attempting to join the domain "ITIA": The specified account already exists. (Note that until a while ago samba wasn't configured to automatically create computer accounts. What I did whenever I wanted an XP to join was to manually create it. When I first attempted to solve the Windows 7 join problem, I setup samba to do this automatically, as this is what most people do, as I understand, and I thought that it might be related. I haven't attempted to add an XP since I made this change, so I don't know if it works, but whether it works or not, the problem remains.) Update 1: Here are the relevant parts of smb.conf: [global] panic action = /usr/share/samba/panic-action %d workgroup = ITIA server string = Itia file server announce as = NT interfaces = 147.102.160.1 volume = %h passdb backend = ldapsam:ldap://ldap.itia.ntua.gr:389 ldap admin dn = uid=samba,ou=daemons,dc=itia,dc=ntua,dc=gr ldap ssl = off ldap suffix = dc=itia,dc=ntua,dc=gr ldap user suffix = ou=people ldap group suffix = ou=groups ldap machine suffix = ou=computers unix password sync = no add machine script = smbldap-useradd -w -i %u log file = /var/log/samba/samba-log.all log level = 3 max log size = 5000 syslog = 2 socket options = SO_KEEPALIVE TCP_NODELAY encrypt passwords = true password level = 1 security = user domain master = yes local master = no wins support = yes domain logons = yes idmap gid = 1000-2000 Update 2: The server has a single network interface eth1 (also an unused eth0 that shows up only in the kernel boot messages) and two ip addresses; the main, 147.102.160.1, and an additional one, 147.102.160.37, that comes up with "ip addr add 147.102.160.37/32 dev eth1" (used only for a web site that has a different certificate than other web sites served from the same machine). One of the problems I recently faced was that samba was using the latter IP address. I fixed that by adding the "interfaces = 147.102.160.1" statement in smb.conf. Now: acheloos:/etc/apache2# tcpdump host 147.102.160.40 and not port 5900 tcpdump: verbose output suppressed, use -v or -vv for full protocol decode listening on eth1, link-type EN10MB (Ethernet), capture size 65535 bytes 13:13:56.549048 IP lithaios.itia.civil.ntua.gr.netbios-dgm > 147.102.160.255.netbios-dgm: NBT UDP PACKET(138) 13:13:56.549056 ARP, Request who-has acheloos2.itia.civil.ntua.gr tell lithaios.itia.civil.ntua.gr, length 46 13:13:56.549091 ARP, Reply acheloos2.itia.civil.ntua.gr is-at 00:10:4b:b4:9e:59 (oui Unknown), length 28 13:13:56.549324 IP acheloos.itia.civil.ntua.gr.netbios-dgm > lithaios.itia.civil.ntua.gr.netbios-dgm: NBT UDP PACKET(138) 13:13:56.549608 IP lithaios.itia.civil.ntua.gr.netbios-dgm > acheloos2.itia.civil.ntua.gr.netbios-dgm: NBT UDP PACKET(138) 13:13:56.549741 IP acheloos.itia.civil.ntua.gr.netbios-dgm > lithaios.itia.civil.ntua.gr.netbios-dgm: NBT UDP PACKET(138) 13:13:56.550364 IP lithaios.itia.civil.ntua.gr.netbios-dgm > acheloos.itia.civil.ntua.gr.netbios-dgm: NBT UDP PACKET(138) 13:13:56.550468 IP acheloos.itia.civil.ntua.gr.netbios-dgm > lithaios.itia.civil.ntua.gr.netbios-dgm: NBT UDP PACKET(138) (acheloos2 is the second IP address, 147.102.160.37). The above dump occurs when I click "OK" (to join the domain), until it asks me for the username and password of a user that can join the domain. I don't know why the client is contacting the second IP address. I tried temporarily deactivating it, but I still had some related ARP traffic (though I think not IP traffic).

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  • Farseer tutorial for the absolute beginners

    - by Bil Simser
    This post is inspired (and somewhat a direct copy) of a couple of posts Emanuele Feronato wrote back in 2009 about Box2D (his tutorial was ActionScript 3 based for Box2D, this is C# XNA for the Farseer Physics Engine). Here’s what we’re building: What is Farseer The Farseer Physics Engine is a collision detection system with realistic physics responses to help you easily create simple hobby games or complex simulation systems. Farseer was built as a .NET version of Box2D (based on the Box2D.XNA port of Box2D). While the constructs and syntax has changed over the years, the principles remain the same. This tutorial will walk you through exactly what Emanuele create for Flash but we’ll be doing it using C#, XNA and the Windows Phone platform. The first step is to download the library from its home on CodePlex. If you have NuGet installed, you can install the library itself using the NuGet package that but we’ll also be using some code from the Samples source that can only be obtained by downloading the library. Once you download and unpacked the zip file into a folder and open the solution, this is what you will get: The Samples XNA WP7 project (and content) have all the demos for Farseer. There’s a wealth of info here and great examples to look at to learn. The Farseer Physics XNA WP7 project contains the core libraries that do all the work. DebugView XNA contains an XNA-ready class to let you view debug data and information in the game draw loop (which you can copy into your project or build the source and reference the assembly). The downloaded version has to be compiled as it’s only available in source format so you can do that now if you want (open the solution file and rebuild everything). If you’re using the NuGet package you can just install that. We only need the core library and we’ll be copying in some code from the samples later. Your first Farseer experiment Start Visual Studio and create a new project using the Windows Phone template can call it whatever you want. It’s time to edit Game1.cs 1 public class Game1 : Game 2 { 3 private readonly GraphicsDeviceManager _graphics; 4 private DebugViewXNA _debugView; 5 private Body _floor; 6 private SpriteBatch _spriteBatch; 7 private float _timer; 8 private World _world; 9 10 public Game1() 11 { 12 _graphics = new GraphicsDeviceManager(this) 13 { 14 PreferredBackBufferHeight = 800, 15 PreferredBackBufferWidth = 480, 16 IsFullScreen = true 17 }; 18 19 Content.RootDirectory = "Content"; 20 21 // Frame rate is 30 fps by default for Windows Phone. 22 TargetElapsedTime = TimeSpan.FromTicks(333333); 23 24 // Extend battery life under lock. 25 InactiveSleepTime = TimeSpan.FromSeconds(1); 26 } 27 28 protected override void LoadContent() 29 { 30 // Create a new SpriteBatch, which can be used to draw textures. 31 _spriteBatch = new SpriteBatch(_graphics.GraphicsDevice); 32 33 // Load our font (DebugViewXNA needs it for the DebugPanel) 34 Content.Load<SpriteFont>("font"); 35 36 // Create our World with a gravity of 10 vertical units 37 if (_world == null) 38 { 39 _world = new World(Vector2.UnitY*10); 40 } 41 else 42 { 43 _world.Clear(); 44 } 45 46 if (_debugView == null) 47 { 48 _debugView = new DebugViewXNA(_world); 49 50 // default is shape, controller, joints 51 // we just want shapes to display 52 _debugView.RemoveFlags(DebugViewFlags.Controllers); 53 _debugView.RemoveFlags(DebugViewFlags.Joint); 54 55 _debugView.LoadContent(GraphicsDevice, Content); 56 } 57 58 // Create and position our floor 59 _floor = BodyFactory.CreateRectangle( 60 _world, 61 ConvertUnits.ToSimUnits(480), 62 ConvertUnits.ToSimUnits(50), 63 10f); 64 _floor.Position = ConvertUnits.ToSimUnits(240, 775); 65 _floor.IsStatic = true; 66 _floor.Restitution = 0.2f; 67 _floor.Friction = 0.2f; 68 } 69 70 protected override void Update(GameTime gameTime) 71 { 72 // Allows the game to exit 73 if (GamePad.GetState(PlayerIndex.One).Buttons.Back == ButtonState.Pressed) 74 Exit(); 75 76 // Create a random box every second 77 _timer += (float) gameTime.ElapsedGameTime.TotalSeconds; 78 if (_timer >= 1.0f) 79 { 80 // Reset our timer 81 _timer = 0f; 82 83 // Determine a random size for each box 84 var random = new Random(); 85 var width = random.Next(20, 100); 86 var height = random.Next(20, 100); 87 88 // Create it and store the size in the user data 89 var box = BodyFactory.CreateRectangle( 90 _world, 91 ConvertUnits.ToSimUnits(width), 92 ConvertUnits.ToSimUnits(height), 93 10f, 94 new Point(width, height)); 95 96 box.BodyType = BodyType.Dynamic; 97 box.Restitution = 0.2f; 98 box.Friction = 0.2f; 99 100 // Randomly pick a location along the top to drop it from 101 box.Position = ConvertUnits.ToSimUnits(random.Next(50, 400), 0); 102 } 103 104 // Advance all the elements in the world 105 _world.Step(Math.Min((float) gameTime.ElapsedGameTime.TotalMilliseconds*0.001f, (1f/30f))); 106 107 // Clean up any boxes that have fallen offscreen 108 foreach (var box in from box in _world.BodyList 109 let pos = ConvertUnits.ToDisplayUnits(box.Position) 110 where pos.Y > _graphics.GraphicsDevice.Viewport.Height 111 select box) 112 { 113 _world.RemoveBody(box); 114 } 115 116 base.Update(gameTime); 117 } 118 119 protected override void Draw(GameTime gameTime) 120 { 121 GraphicsDevice.Clear(Color.FromNonPremultiplied(51, 51, 51, 255)); 122 123 _spriteBatch.Begin(); 124 125 var projection = Matrix.CreateOrthographicOffCenter( 126 0f, 127 ConvertUnits.ToSimUnits(_graphics.GraphicsDevice.Viewport.Width), 128 ConvertUnits.ToSimUnits(_graphics.GraphicsDevice.Viewport.Height), 0f, 0f, 129 1f); 130 _debugView.RenderDebugData(ref projection); 131 132 _spriteBatch.End(); 133 134 base.Draw(gameTime); 135 } 136 } 137 Lines 4: Declare the debug view we’ll use for rendering (more on that later). Lines 8: Declare _world variable of type class World. World is the main object to interact with the Farseer engine. It stores all the joints and bodies, and is responsible for stepping through the simulation. Lines 12-17: Create the graphics device we’ll be rendering on. This is an XNA component and we’re just setting it to be the same size as the phone and toggling it to be full screen (no system tray). Lines 34: We create a SpriteFont here by adding it to the project. It’s called “font” because that’s what the DebugView uses but you can name it whatever you want (and if you’re not using DebugView for your production app you might have several fonts). Lines 37-44: We create the physics environment that Farseer uses to contain all the objects by specifying it here. We’re using Vector2.UnitY*10 to represent the gravity to be used in the environment. In other words, 10 units going in a downward motion. Lines 46-56: We create the DebugViewXNA here. This is copied from the […] from the code you downloaded and provides the ability to render all entities onto the screen. In a production release you’ll be doing the rendering yourself of each object but we cheat a bit for the demo and let the DebugView do it for us. The other thing it can provide is to render out a panel of debugging information while the simulation is going on. This is useful in tracking down objects, figuring out how something works, or just keeping track of what’s in the engine. Lines 49-67: Here we create a rigid body (Farseer only supports rigid bodies) to represent the floor that we’ll drop objects onto. We create it by using one of the Farseer factories and specifying the width and height. The ConvertUnits class is copied from the samples code as-is and lets us toggle between display units (pixels) and simulation units (usually metres). We’re creating a floor that’s 480 pixels wide and 50 pixels high (converting them to SimUnits for the engine to understand). We also position it near the bottom of the screen. Values are in metres and when specifying values they refer to the centre of the body object. Lines 77-78: The game Update method fires 30 times a second, too fast to be creating objects this quickly. So we use a variable to track the elapsed seconds since the last update, accumulate that value, then create a new box to drop when 1 second has passed. Lines 89-94: We create a box the same way we created our floor (coming up with a random width and height for the box). Lines 96-101: We set the box to be Dynamic (rather than Static like the floor object) and position it somewhere along the top of the screen. And now you created the world. Gravity does the rest and the boxes fall to the ground. Here’s the result: Farseer Physics Engine Demo using XNA Lines 105: We must update the world at every frame. We do this with the Step method which takes in the time interval. [more] Lines 108-114: Body objects are added to the world but never automatically removed (because Farseer doesn’t know about the display world, it has no idea if an item is on the screen or not). Here we just loop through all the entities and anything that’s dropped off the screen (below the bottom) gets removed from the World. This keeps our entity count down (the simulation never has more than 30 or 40 objects in the world no matter how long you run it for). Too many entities and the app will grind to a halt. Lines 125-130: Farseer knows nothing about the UI so that’s entirely up to you as to how to draw things. Farseer is just tracking the objects and moving them around using the physics engine and it’s rules. You’ll still use XNA to draw items (using the SpriteBatch.Draw method) so you can load up your usual textures and draw items and pirates and dancing zombies all over the screen. Instead in this demo we’re going to cheat a little. In the sample code for Farseer you can download there’s a project called DebugView XNA. This project contains the DebugViewXNA class which just handles iterating through all the bodies in the world and drawing the shapes. So we call the RenderDebugData method here of that class to draw everything correctly. In the case of this demo, we just want to draw Shapes so take a look at the source code for the DebugViewXNA class as to how it extracts all the vertices for the shapes created (in this case simple boxes) and draws them. You’ll learn a *lot* about how Farseer works just by looking at this class. That’s it, that’s all. Simple huh? Hope you enjoy the code and library. Physics is hard and requires some math skills to really grok. The Farseer Physics Engine makes it pretty easy to get up and running and start building games. In future posts we’ll get more in-depth with things you can do with the engine so this is just the beginning. Enjoy!

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  • Taming Hopping Windows

    - by Roman Schindlauer
    At first glance, hopping windows seem fairly innocuous and obvious. They organize events into windows with a simple periodic definition: the windows have some duration d (e.g. a window covers 5 second time intervals), an interval or period p (e.g. a new window starts every 2 seconds) and an alignment a (e.g. one of those windows starts at 12:00 PM on March 15, 2012 UTC). var wins = xs     .HoppingWindow(TimeSpan.FromSeconds(5),                    TimeSpan.FromSeconds(2),                    new DateTime(2012, 3, 15, 12, 0, 0, DateTimeKind.Utc)); Logically, there is a window with start time a + np and end time a + np + d for every integer n. That’s a lot of windows. So why doesn’t the following query (always) blow up? var query = wins.Select(win => win.Count()); A few users have asked why StreamInsight doesn’t produce output for empty windows. Primarily it’s because there is an infinite number of empty windows! (Actually, StreamInsight uses DateTimeOffset.MaxValue to approximate “the end of time” and DateTimeOffset.MinValue to approximate “the beginning of time”, so the number of windows is lower in practice.) That was the good news. Now the bad news. Events also have duration. Consider the following simple input: var xs = this.Application                 .DefineEnumerable(() => new[]                     { EdgeEvent.CreateStart(DateTimeOffset.UtcNow, 0) })                 .ToStreamable(AdvanceTimeSettings.IncreasingStartTime); Because the event has no explicit end edge, it lasts until the end of time. So there are lots of non-empty windows if we apply a hopping window to that single event! For this reason, we need to be careful with hopping window queries in StreamInsight. Or we can switch to a custom implementation of hopping windows that doesn’t suffer from this shortcoming. The alternate window implementation produces output only when the input changes. We start by breaking up the timeline into non-overlapping intervals assigned to each window. In figure 1, six hopping windows (“Windows”) are assigned to six intervals (“Assignments”) in the timeline. Next we take input events (“Events”) and alter their lifetimes (“Altered Events”) so that they cover the intervals of the windows they intersect. In figure 1, you can see that the first event e1 intersects windows w1 and w2 so it is adjusted to cover assignments a1 and a2. Finally, we can use snapshot windows (“Snapshots”) to produce output for the hopping windows. Notice however that instead of having six windows generating output, we have only four. The first and second snapshots correspond to the first and second hopping windows. The remaining snapshots however cover two hopping windows each! While in this example we saved only two events, the savings can be more significant when the ratio of event duration to window duration is higher. Figure 1: Timeline The implementation of this strategy is straightforward. We need to set the start times of events to the start time of the interval assigned to the earliest window including the start time. Similarly, we need to modify the end times of events to the end time of the interval assigned to the latest window including the end time. The following snap-to-boundary function that rounds a timestamp value t down to the nearest value t' <= t such that t' is a + np for some integer n will be useful. For convenience, we will represent both DateTime and TimeSpan values using long ticks: static long SnapToBoundary(long t, long a, long p) {     return t - ((t - a) % p) - (t > a ? 0L : p); } How do we find the earliest window including the start time for an event? It’s the window following the last window that does not include the start time assuming that there are no gaps in the windows (i.e. duration < interval), and limitation of this solution. To find the end time of that antecedent window, we need to know the alignment of window ends: long e = a + (d % p); Using the window end alignment, we are finally ready to describe the start time selector: static long AdjustStartTime(long t, long e, long p) {     return SnapToBoundary(t, e, p) + p; } To find the latest window including the end time for an event, we look for the last window start time (non-inclusive): public static long AdjustEndTime(long t, long a, long d, long p) {     return SnapToBoundary(t - 1, a, p) + p + d; } Bringing it together, we can define the translation from events to ‘altered events’ as in Figure 1: public static IQStreamable<T> SnapToWindowIntervals<T>(IQStreamable<T> source, TimeSpan duration, TimeSpan interval, DateTime alignment) {     if (source == null) throw new ArgumentNullException("source");     // reason about DateTime and TimeSpan in ticks     long d = Math.Min(DateTime.MaxValue.Ticks, duration.Ticks);     long p = Math.Min(DateTime.MaxValue.Ticks, Math.Abs(interval.Ticks));     // set alignment to earliest possible window     var a = alignment.ToUniversalTime().Ticks % p;     // verify constraints of this solution     if (d <= 0L) { throw new ArgumentOutOfRangeException("duration"); }     if (p == 0L || p > d) { throw new ArgumentOutOfRangeException("interval"); }     // find the alignment of window ends     long e = a + (d % p);     return source.AlterEventLifetime(         evt => ToDateTime(AdjustStartTime(evt.StartTime.ToUniversalTime().Ticks, e, p)),         evt => ToDateTime(AdjustEndTime(evt.EndTime.ToUniversalTime().Ticks, a, d, p)) -             ToDateTime(AdjustStartTime(evt.StartTime.ToUniversalTime().Ticks, e, p))); } public static DateTime ToDateTime(long ticks) {     // just snap to min or max value rather than under/overflowing     return ticks < DateTime.MinValue.Ticks         ? new DateTime(DateTime.MinValue.Ticks, DateTimeKind.Utc)         : ticks > DateTime.MaxValue.Ticks         ? new DateTime(DateTime.MaxValue.Ticks, DateTimeKind.Utc)         : new DateTime(ticks, DateTimeKind.Utc); } Finally, we can describe our custom hopping window operator: public static IQWindowedStreamable<T> HoppingWindow2<T>(     IQStreamable<T> source,     TimeSpan duration,     TimeSpan interval,     DateTime alignment) {     if (source == null) { throw new ArgumentNullException("source"); }     return SnapToWindowIntervals(source, duration, interval, alignment).SnapshotWindow(); } By switching from HoppingWindow to HoppingWindow2 in the following example, the query returns quickly rather than gobbling resources and ultimately failing! public void Main() {     var start = new DateTimeOffset(new DateTime(2012, 6, 28), TimeSpan.Zero);     var duration = TimeSpan.FromSeconds(5);     var interval = TimeSpan.FromSeconds(2);     var alignment = new DateTime(2012, 3, 15, 12, 0, 0, DateTimeKind.Utc);     var events = this.Application.DefineEnumerable(() => new[]     {         EdgeEvent.CreateStart(start.AddSeconds(0), "e0"),         EdgeEvent.CreateStart(start.AddSeconds(1), "e1"),         EdgeEvent.CreateEnd(start.AddSeconds(1), start.AddSeconds(2), "e1"),         EdgeEvent.CreateStart(start.AddSeconds(3), "e2"),         EdgeEvent.CreateStart(start.AddSeconds(9), "e3"),         EdgeEvent.CreateEnd(start.AddSeconds(3), start.AddSeconds(10), "e2"),         EdgeEvent.CreateEnd(start.AddSeconds(9), start.AddSeconds(10), "e3"),     }).ToStreamable(AdvanceTimeSettings.IncreasingStartTime);     var adjustedEvents = SnapToWindowIntervals(events, duration, interval, alignment);     var query = from win in HoppingWindow2(events, duration, interval, alignment)                 select win.Count();     DisplayResults(adjustedEvents, "Adjusted Events");     DisplayResults(query, "Query"); } As you can see, instead of producing a massive number of windows for the open start edge e0, a single window is emitted from 12:00:15 AM until the end of time: Adjusted Events StartTime EndTime Payload 6/28/2012 12:00:01 AM 12/31/9999 11:59:59 PM e0 6/28/2012 12:00:03 AM 6/28/2012 12:00:07 AM e1 6/28/2012 12:00:05 AM 6/28/2012 12:00:15 AM e2 6/28/2012 12:00:11 AM 6/28/2012 12:00:15 AM e3 Query StartTime EndTime Payload 6/28/2012 12:00:01 AM 6/28/2012 12:00:03 AM 1 6/28/2012 12:00:03 AM 6/28/2012 12:00:05 AM 2 6/28/2012 12:00:05 AM 6/28/2012 12:00:07 AM 3 6/28/2012 12:00:07 AM 6/28/2012 12:00:11 AM 2 6/28/2012 12:00:11 AM 6/28/2012 12:00:15 AM 3 6/28/2012 12:00:15 AM 12/31/9999 11:59:59 PM 1 Regards, The StreamInsight Team

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  • Project Jigsaw: Late for the train: The Q&A

    - by Mark Reinhold
    I recently proposed, to the Java community in general and to the SE 8 (JSR 337) Expert Group in particular, to defer Project Jigsaw from Java 8 to Java 9. I also proposed to aim explicitly for a regular two-year release cycle going forward. Herewith a summary of the key questions I’ve seen in reaction to these proposals, along with answers. Making the decision Q Has the Java SE 8 Expert Group decided whether to defer the addition of a module system and the modularization of the Platform to Java SE 9? A No, it has not yet decided. Q By when do you expect the EG to make this decision? A In the next month or so. Q How can I make sure my voice is heard? A The EG will consider all relevant input from the wider community. If you have a prominent blog, column, or other communication channel then there’s a good chance that we’ve already seen your opinion. If not, you’re welcome to send it to the Java SE 8 Comments List, which is the EG’s official feedback channel. Q What’s the overall tone of the feedback you’ve received? A The feedback has been about evenly divided as to whether Java 8 should be delayed for Jigsaw, Jigsaw should be deferred to Java 9, or some other, usually less-realistic, option should be taken. Project Jigsaw Q Why is Project Jigsaw taking so long? A Project Jigsaw started at Sun, way back in August 2008. Like many efforts during the final years of Sun, it was not well staffed. Jigsaw initially ran on a shoestring, with just a handful of mostly part-time engineers, so progress was slow. During the integration of Sun into Oracle all work on Jigsaw was halted for a time, but it was eventually resumed after a thorough consideration of the alternatives. Project Jigsaw was really only fully staffed about a year ago, around the time that Java 7 shipped. We’ve added a few more engineers to the team since then, but that can’t make up for the inadequate initial staffing and the time lost during the transition. Q So it’s really just a matter of staffing limitations and corporate-integration distractions? A Aside from these difficulties, the other main factor in the duration of the project is the sheer technical difficulty of modularizing the JDK. Q Why is modularizing the JDK so hard? A There are two main reasons. The first is that the JDK code base is deeply interconnected at both the API and the implementation levels, having been built over many years primarily in the style of a monolithic software system. We’ve spent considerable effort eliminating or at least simplifying as many API and implementation dependences as possible, so that both the Platform and its implementations can be presented as a coherent set of interdependent modules, but some particularly thorny cases remain. Q What’s the second reason? A We want to maintain as much compatibility with prior releases as possible, most especially for existing classpath-based applications but also, to the extent feasible, for applications composed of modules. Q Is modularizing the JDK even necessary? Can’t you just put it in one big module? A Modularizing the JDK, and more specifically modularizing the Java SE Platform, will enable standard yet flexible Java runtime configurations scaling from large servers down to small embedded devices. In the long term it will enable the convergence of Java SE with the higher-end Java ME Platforms. Q Is Project Jigsaw just about modularizing the JDK? A As originally conceived, Project Jigsaw was indeed focused primarily upon modularizing the JDK. The growing demand for a truly standard module system for the Java Platform, which could be used not just for the Platform itself but also for libraries and applications built on top of it, later motivated expanding the scope of the effort. Q As a developer, why should I care about Project Jigsaw? A The introduction of a modular Java Platform will, in the long term, fundamentally change the way that Java implementations, libraries, frameworks, tools, and applications are designed, built, and deployed. Q How much progress has Project Jigsaw made? A We’ve actually made a lot of progress. Much of the core functionality of the module system has been prototyped and works at both compile time and run time. We’ve extended the Java programming language with module declarations, worked out a structure for modular source trees and corresponding compiled-class trees, and implemented these features in javac. We’ve defined an efficient module-file format, extended the JVM to bootstrap a modular JRE, and designed and implemented a preliminary API. We’ve used the module system to make a good first cut at dividing the JDK and the Java SE API into a coherent set of modules. Among other things, we’re currently working to retrofit the java.util.ServiceLoader API to support modular services. Q I want to help! How can I get involved? A Check out the project page, read the draft requirements and design overview documents, download the latest prototype build, and play with it. You can tell us what you think, and follow the rest of our work in real time, on the jigsaw-dev list. The Java Platform Module System JSR Q What’s the relationship between Project Jigsaw and the eventual Java Platform Module System JSR? A At a high level, Project Jigsaw has two phases. In the first phase we’re exploring an approach to modularity that’s markedly different from that of existing Java modularity solutions. We’ve assumed that we can change the Java programming language, the virtual machine, and the APIs. Doing so enables a design which can strongly enforce module boundaries in all program phases, from compilation to deployment to execution. That, in turn, leads to better usability, diagnosability, security, and performance. The ultimate goal of the first phase is produce a working prototype which can inform the work of the Module-System JSR EG. Q What will happen in the second phase of Project Jigsaw? A The second phase will produce the reference implementation of the specification created by the Module-System JSR EG. The EG might ultimately choose an entirely different approach than the one we’re exploring now. If and when that happens then Project Jigsaw will change course as necessary, but either way I think that the end result will be better for having been informed by our current work. Maven & OSGi Q Why not just use Maven? A Maven is a software project management and comprehension tool. As such it can be seen as a kind of build-time module system but, by its nature, it does nothing to support modularity at run time. Q Why not just adopt OSGi? A OSGi is a rich dynamic component system which includes not just a module system but also a life-cycle model and a dynamic service registry. The latter two facilities are useful to some kinds of sophisticated applications, but I don’t think they’re of wide enough interest to be standardized as part of the Java SE Platform. Q Okay, then why not just adopt the module layer of OSGi? A The OSGi module layer is not operative at compile time; it only addresses modularity during packaging, deployment, and execution. As it stands, moreover, it’s useful for library and application modules but, since it’s built strictly on top of the Java SE Platform, it can’t be used to modularize the Platform itself. Q If Maven addresses modularity at build time, and the OSGi module layer addresses modularity during deployment and at run time, then why not just use the two together, as many developers already do? A The combination of Maven and OSGi is certainly very useful in practice today. These systems have, however, been built on top of the existing Java platform; they have not been able to change the platform itself. This means, among other things, that module boundaries are weakly enforced, if at all, which makes it difficult to diagnose configuration errors and impossible to run untrusted code securely. The prototype Jigsaw module system, by contrast, aims to define a platform-level solution which extends both the language and the JVM in order to enforce module boundaries strongly and uniformly in all program phases. Q If the EG chooses an approach like the one currently being taken in the Jigsaw prototype, will Maven and OSGi be made obsolete? A No, not at all! No matter what approach is taken, to ensure wide adoption it’s essential that the standard Java Platform Module System interact well with Maven. Applications that depend upon the sophisticated features of OSGi will no doubt continue to use OSGi, so it’s critical that implementations of OSGi be able to run on top of the Java module system and, if suitably modified, support OSGi bundles that depend upon Java modules. Ideas for how to do that are currently being explored in Project Penrose. Java 8 & Java 9 Q Without Jigsaw, won’t Java 8 be a pretty boring release? A No, far from it! It’s still slated to include the widely-anticipated Project Lambda (JSR 335), work on which has been going very well, along with the new Date/Time API (JSR 310), Type Annotations (JSR 308), and a set of smaller features already in progress. Q Won’t deferring Jigsaw to Java 9 delay the eventual convergence of the higher-end Java ME Platforms with Java SE? A It will slow that transition, but it will not stop it. To allow progress toward that convergence to be made with Java 8 I’ve suggested to the Java SE 8 EG that we consider specifying a small number of Profiles which would allow compact configurations of the SE Platform to be built and deployed. Q If Jigsaw is deferred to Java 9, would the Oracle engineers currently working on it be reassigned to other Java 8 features and then return to working on Jigsaw again after Java 8 ships? A No, these engineers would continue to work primarily on Jigsaw from now until Java 9 ships. Q Why not drop Lambda and finish Jigsaw instead? A Even if the engineers currently working on Lambda could instantly switch over to Jigsaw and immediately become productive—which of course they can’t—there are less than nine months remaining in the Java 8 schedule for work on major features. That’s just not enough time for the broad review, testing, and feedback which such a fundamental change to the Java Platform requires. Q Why not ship the module system in Java 8, and then modularize the platform in Java 9? A If we deliver a module system in one release but don’t use it to modularize the JDK until some later release then we run a big risk of getting something fundamentally wrong. If that happens then we’d have to fix it in the later release, and fixing fundamental design flaws after the fact almost always leads to a poor end result. Q Why not ship Jigsaw in an 8.5 release, less than two years after 8? Or why not just ship a new release every year, rather than every other year? A Many more developers work on the JDK today than a couple of years ago, both because Oracle has dramatically increased its own investment and because other organizations and individuals have joined the OpenJDK Community. Collectively we don’t, however, have the bandwidth required to ship and then provide long-term support for a big JDK release more frequently than about every other year. Q What’s the feedback been on the two-year release-cycle proposal? A For just about every comment that we should release more frequently, so that new features are available sooner, there’s been another asking for an even slower release cycle so that large teams of enterprise developers who ship mission-critical applications have a chance to migrate at a comfortable pace.

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  • Stumbling Through: Visual Studio 2010 (Part IV)

    So finally we get to the fun part the fruits of all of our middle-tier/back end labors of generating classes to interface with an XML data source that the previous posts were about can now be presented quickly and easily to an end user.  I think.  Well see.  Well be using a WPF window to display all of our various MFL information that weve collected in the two XML files, and well provide a means of adding, updating and deleting each of these entities using as little code as possible.  Additionally, I would like to dig into the performance of this solution as well as the flexibility of it if were were to modify the underlying XML schema.  So first things first, lets create a WPF project and include our xml data in a data folder within.  On the main window, well drag out the following controls: A combo box to contain all of the teams A list box to show the players of the selected team, along with add/delete player buttons A text box tied to the selected players name, with a save button to save any changes made to the player name A combo box of all the available positions, tied to the currently selected players position A data grid tied to the statistics of the currently selected player, with add/delete statistic buttons This monstrosity of a form and its associated project will look like this (dont forget to reference the DataFoundation project from the Presentation project): To get to the visual data binding, as we learned in a previous post, you have to first make sure the project containing your bindable classes is compiled.  Do so, and then open the Data Sources pane to add a reference to the Teams and Positions classes in the DataFoundation project: Why only Team and Position?  Well, we will get to Players from Teams, and Statistics from Players so no need to make an interface for them as well see in a second.  As for Positions, well need a way to bind the dropdown to ALL positions they dont appear underneath any of the other classes so we need to reference it directly.  After adding these guys, expand every node in your Data Sources pane and see how the Team node allows you to drill into Players and then Statistics.  This is why there was no need to bring in a reference to those classes for the UI we are designing: Now for the seriously hard work of binding all of our controls to the correct data sources.  Drag the following items from the Data Sources pane to the specified control on the window design canvas: Team.Name > Teams combo box Team.Players.Name > Players list box Team.Players.Name > Player name text box Team.Players.Statistics > Statistics data grid Position.Name > Positions combo box That is it!  Really?  Well, no, not really there is one caveat here in that the Positions combo box is not bound the selected players position.  To do so, we will apply a binding to the position combo boxs SelectedValue to point to the current players PositionId value: That should do the trick now, all we need to worry about is loading the actual data.  Sadly, it appears as if we will need to drop to code in order to invoke our IO methods to load all teams and positions.  At least Visual Studio kindly created the stubs for us to do so, ultimately the code should look like this: Note the weirdness with the InitializeDataFiles call that is my current means of telling an IO where to load the data for each of the entities.  I havent thought of a more intuitive way than that yet, but do note that all data is loaded from Teams.xml besides for positions, which is loaded from Lookups.xml.   I think that may be all we need to do to at least load all of the data, lets run it and see: Yay!  All of our glorious data is being displayed!  Er, wait, whats up with the position dropdown?  Why is it red?  Lets select the RB and see if everything updates: Crap, the position didnt update to reflect the selected player, but everything else did.  Where did we go wrong in binding the position to the selected player?  Thinking about it a bit and comparing it to how traditional data binding works, I realize that we never set the value member (or some similar property) to tell the control to join the Id of the source (positions) to the position Id of the player.  I dont see a similar property to that on the combo box control, but I do see a property named SelectedValuePath that might be it, so I set it to Id and run the app again: Hey, all right!  No red box around the positions combo box.  Unfortunately, selecting the RB does not update the dropdown to point to Runningback.  Hmmm.  Now what could it be?  Maybe the problem is that we are loading teams before we are loading positions, so when it binds position Id, all of the positions arent loaded yet.  I went to the code behind and switched things so position loads first and no dice.  Same result when I run.  Why?  WHY?  Ok, ok, calm down, take a deep breath.  Get something with caffeine or sugar (preferably both) and think rationally. Ok, gigantic chocolate chip cookie and a mountain dew chaser have never let me down in the past, so dont fail me now!  Ah ha!  of course!  I didnt even have to finish the mountain dew and I think Ive got it:  Data Context.  By default, when setting on the selected value binding for the dropdown, the data context was list_team.  I dont even know what the heck list_team is, we want it to be bound to our team players view source resource instead, like this: Running it now and selecting the various players: Done and done.  Everything read and bound, thank you caffeine and sugar!  Oh, and thank you Visual Studio 2010.  Lets wire up some of those buttons now There has got to be a better way to do this, but it works for now.  What the add player button does is add a new player object to the currently selected team.  Unfortunately, I couldnt get the new object to automatically show up in the players list (something about not using an observable collection gotta look into this) so I just save the change immediately and reload the screen.  Terrible, but it works: Lets go after something easier:  The save button.  By default, as we type in new text for the players name, it is showing up in the list box as updated.  Cool!  Why couldnt my add new player logic do that?  Anyway, the save button should be as simple as invoking MFL.IO.Save for the selected player, like this: MFL.IO.Save((MFL.Player)lbTeamPlayers.SelectedItem, true); Surprisingly, that worked on the first try.  Lets see if we get as lucky with the Delete player button: MFL.IO.Delete((MFL.Player)lbTeamPlayers.SelectedItem); Refresh(); Note the use of the Refresh method again I cant seem to figure out why updates to the underlying data source are immediately reflected, but adds and deletes are not.  That is a problem for another day, and again my hunch is that I should be binding to something more complex than IEnumerable (like observable collection). Now that an example of the basic CRUD methods are wired up, I want to quickly investigate the performance of this beast.  Im going to make a special button to add 30 teams, each with 50 players and 10 seasons worth of stats.  If my math is right, that will end up with 15000 rows of data, a pretty hefty amount for an XML file.  The save of all this new data took a little over a minute, but that is acceptable because we wouldnt typically be saving batches of 15k records, and the resulting XML file size is a little over a megabyte.  Not huge, but big enough to see some read performance numbers or so I thought.  It reads this file and renders the first team in under a second.  That is unbelievable, but we are lazy loading and the file really wasnt that big.  I will increase it to 50 teams with 100 players and 20 seasons each - 100,000 rows.  It took a year and a half to save all of that data, and resulted in an 8 megabyte file.  Seriously, if you are loading XML files this large, get a freaking database!  Despite this, it STILL takes under a second to load and render the first team, which is interesting mostly because I thought that it was loading that entire 8 MB XML file behind the scenes.  I have to say that I am quite impressed with the performance of the LINQ to XML approach, particularly since I took no efforts to optimize any of this code and was fairly new to the concept from the start.  There might be some merit to this little project after all Look out SQL Server and Oracle, use XML files instead!  Next up, I am going to completely pull the rug out from under the UI and change a number of entities in our model.  How well will the code be regenerated?  How much effort will be required to tie things back together in the UI?Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • T60 Screen/LCD gets black after some minutes with a highpitched sound rising and fading

    - by edelwater
    Just now my T60 screen got "black" (so no display). On my second monitor: no problems so the VGA output works. Symptom: Screen blanks / no display, but it works on the second monitor Steps to reproduce: - boot - wait (it does not matter what you do you do not have to login or anything) - (now the monitor of the laptop slowly begins to make a ssssssssHHHHHHHHHHHHHHHHHWOEOEssssssss noise of about 10 seconds) - right after the sounds ends, the monitor gets black. Sometimes it seems to be the same each time. Software: Installed no new software before/after, running ZoneAlarm and antivirus. Other: It does not feel hot in any place, there don't seem to be running processes with strange behaviour. Warranty: Out of warranty What was I doing: Typing text on a website and doing some PHP coding in a text editor. What can I do here other than buy a new laptop? Does it sound familiar to known cases? Update 1: Exactly the same problem: http://forums.lenovo.com/t5/T61-and-prior-T-series-ThinkPad/T60-Screen-Blackout/m-p/288772 and the second poster (garyj), http://forums.lenovo.com/t5/T61-and-prior-T-series-ThinkPad/Black-Screen-on-T60/m-p/235053#M48627 And here: "I have that same problem. I replaced the CCRL on mine and it works fine when the screen is not screwed in. Once the frame of the LCD screen (metal portion) touches the metal on the laptop which holds the screen the screen goes black. If the metal is touching the screen when you boot up it boots up with it being very dimmly lit. " from http://forums.lenovo.com/t5/T61-and-prior-T-series-ThinkPad/T60-screen-problems/m-p/205047#M44995 (it seems replacing the LCD display is no use, he tried it three times). Same problem: http://forums.lenovo.com/t5/T61-and-prior-T-series-ThinkPad/T60-black-screen/m-p/80604#M25914 Hmmm... not handy 3 or 4 months ago I ordered and installed a new fan. Now the LCD. Which does not seem the core issue but some electric issue so it seems replacing the LCD is not the thing to do here. If it is not the LCD that needs to be replaced (see other threads), which parts can I order to fix this? Is there any information which could lead me to identify the issue? I have read replacing the "inverter" AND the "backlightning" would that make sense? Update 2: I replaced the inverter with another inverter, but IO have the same problem. I DID notice that the inverter is the component that makes the sssssssssssssHHHHHHHHHH sound AND it becomes very hot in a few seconds. (So both the old and the test one) The problem is hmmm wat is then the thing that makes the inverter hot by (assumption) after which it shuts itself down. Is it either the input or the output? The output seems to me not, because the screen seems to function so it must be the electricity coming in. But what causes it to become so hot would it be the VGA card outputting some unusual high voltage seems unlikely? I am looking for the component to order / replace update 3: Great news. Ewendish gave me the hint to look in the BIOS. While I was in the BIOS I noticed that the screen did not switch off and there was not a high pitched sound. So I lowered some settings in the BIOS. I then noticed that with brightness turned to 0 (via FN End), it does not make a high pitched sound and does not turn off, with brightness turned up just three "stripes" it starts making the sound. So I could from now on work under lowest brightness modus or... see where the problem lies. So as stated below with either power management or display drivers / ATI Catalyst settings / Windows display settings. I'm trying to see where it lies, but I will google some first. Update 4: I wiped clean the Windows XP installation and installed Windows 7 on it. Unfortunately the problem remains: as soon as the brightness goes up the screen starts hissing. This means... back to original thought: it probably IS a hardware problem. Although ... again... if it is NOT the inverter, what is it? Could it be the backlightning component? I could try to switch that with a another T60... but this is quite tricky.

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  • Odd tcp deadlock under windows

    - by John Robertson
    We are moving large amounts of data on a LAN and it has to happen very rapidly and reliably. Currently we use windows TCP as implemented in C++. Using large (synchronous) sends moves the data much faster than a bunch of smaller (synchronous) sends but will frequently deadlock for large gaps of time (.15 seconds) causing the overall transfer rate to plummet. This deadlock happens in very particular circumstances which makes me believe it should be preventable altogether. More importantly if we don't really know the cause we don't really know it won't happen some time with smaller sends anyway. Can anyone explain this deadlock? Deadlock description (OK, zombie-locked, it isn't dead, but for .15 or so seconds it stops, then starts again) The receiving side sends an ACK. The sending side sends a packet containing the end of a message (push flag is set) The call to socket.recv takes about .15 seconds(!) to return About the time the call returns an ACK is sent by the receiving side The the next packet from the sender is finally sent (why is it waiting? the tcp window is plenty big) The odd thing about (3) is that typically that call doesn't take much time at all and receives exactly the same amount of data. On a 2Ghz machine that's 300 million instructions worth of time. I am assuming the call doesn't (heaven forbid) wait for the received data to be acked before it returns, so the ack must be waiting for the call to return, or both must be delayed by something else. The problem NEVER happens when there is a second packet of data (part of the same message) arriving between 1 and 2. That part very clearly makes it sound like it has to do with the fact that windows TCP will not send back a no-data ACK until either a second packet arrives or a 200ms timer expires. However the delay is less than 200 ms (its more like 150 ms). The third unseemly character (and to my mind the real culprit) is (5). Send is definitely being called well before that .15 seconds is up, but the data NEVER hits the wire before that ack returns. That is the most bizarre part of this deadlock to me. Its not a tcp blockage because the TCP window is plenty big since we set SO_RCVBUF to something like 500*1460 (which is still under a meg). The data is coming in very fast (basically there is a loop spinning out data via send) so the buffer should fill almost immediately. According to msdn the buffer being full and at least one pending send should cause the data to be sent (though in another place it mentions that there various "heuristics" used in deciding when a send hits the wire). Anway, why the sender doesn't actually send more data during that .15 second pause is the most bizarre part to me. The information above was captured on the receiving side via wireshark (except of course the socket.recv return times which were logged in a text file). We tried changing the send buffer to zero and turning off Nagle on the sender (yes, I know Nagle is about not sending small packets - but we tried turning Nagle off in case that was part of the unstated "heuristics" affecting whether the message would be posted to the wire. Technically microsoft's Nagle is that a small packet isn't sent if the buffer is full and there is an outstanding ACK, so it seemed like a possibility).

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  • Is this question too hard for a seasoned C++ architect?

    - by Monomer
    Background Information We're looking to hire a seasoned C++ architect (10+years dev, of which at least 6years must be C++ ) for a high frequency trading platform. Job advert says STL, Boost proficiency is a must with preferences to modern uses of C++. The company I work for is a Fortune 500 IB (aka finance industry), it requires passes in all the standard SHL tests (numeric, vocab, spatial etc) before interviews can commence. Everyone on the team was given the task of coming up with one question to ask the candidates during a written/typed test, please note this is the second test provided to the candidates, the first being Advanced IKM C++ test, done in the offices supervised and without internet access. People passing that do the second test. After roughly 70 candidates, my question has been determined to be statistically the worst performing - aka least number of people attempted it, furthermore even less people were able to give meaningful answers. Please note, the second test is not timed, the candidate can literally take as long as they like (we've had one person take roughly 10.5hrs) My question to SO is this, after SHL and IKM adv c++ tests, backed up with at least 6+ years C++ development experience, is it still ok not to be able to even comment about let alone come up with some loose strategy for solving the following question. The Question There is a class C with methods foo, boo, boo_and_foo and foo_and_boo. Each method takes i,j,k and l clock cycles respectively, where i < j, k < i+j and l < i+j. class C { public: int foo() {...} int boo() {...} int boo_and_foo() {...} int foo_and_boo() {...} }; In code one might write: C c; . . int i = c.foo() + c.boo(); But it would be better to have: int i = c.foo_and_boo(); What changes or techniques could one make to the definition of C, that would allow similar syntax of the original usage, but instead have the compiler generate the latter. Note that foo and boo are not commutative. Possible Solution We were basically looking for an expression templates based approach, and were willing to give marks to anyone who had even hinted or used the phrase or related terminology. We got only two people that used the wording, but weren't able to properly describe how they accomplish the task in detail. We use such techniques all over the place, due to the use of various mathematical operators for matrix and vector based calculations, for example to decide when to use IPP or hand woven implementations at compile time for a particular architecture and many other things. The particular area of software development requires microsecond response times. I believe could/should be able to teach a junior such techniques, but given the assumed caliber of candidates I expected a little more. Is this really a difficult question? Should it be removed? Or are we just not seeing the right candidates?

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  • Trouble calculating correct decimal digits.

    - by Crath
    I am trying to create a program that will do some simple calculations, but am having trouble with the program not doing the correct math, or placing the decimal correctly, or something. Some other people I asked cannot figure it out either. Here is the code: http://pastie.org/887352 When you enter the following data: Weekly Wage: 500 Raise: 3 Years Employed: 8 It outputs the following data: Year Annual Salary 1 $26000.00 2 $26780.00 3 $27560.00 4 $28340.00 5 $29120.00 6 $29900.00 7 $30680.00 8 $31460.00 And it should be outputting: Year Annual Salary 1 $26000.00 2 $26780.00 3 $27583.40 4 $28410.90 5 $29263.23 6 $30141.13 7 $31045.36 8 $31976.72 Here is the full description of the task: 8.17 ( Pay Raise Calculator Application) Develop an application that computes the amount of money an employee makes each year over a user- specified number of years. Assume the employee receives a pay raise once every year. The user specifies in the application the initial weekly salary, the amount of the raise (in percent per year) and the number of years for which the amounts earned will be calculated. The application should run as shown in Fig. 8.22. in your text. (fig 8.22 is the output i posted above as what my program should be posting) Opening the template source code file. Open the PayRaise.cpp file in your text editor or IDE. Defining variables and prompting the user for input. To store the raise percentage and years of employment that the user inputs, define int variables rate and years, in main after line 12. Also define double variable wage to store the user’s annual wage. Then, insert statements that prompt the user for the raise percentage, years of employment and starting weekly wage. Store the values typed at the keyboard in the rate, years and wage variables, respectively. To find the annual wage, multiply the new wage by 52 (the number of weeks per year) and store the result in wage. Displaying a table header and formatting output. Use the left and setw stream manipulators to display a table header as shown in Fig. 8.22 in your text. The first column should be six characters wide. Then use the fixed and setprecision stream manipulators to format floating- point values with two positions to the left of the decimal point. Writing a for statement header. Insert a for statement. Before the first semicolon in the for statement header, define and initialize the variable counter to 1. Before the second semicolon, enter a loop- continuation condition that will cause the for statement to loop until counter has reached the number of years entered. After the second semicolon, enter the increment of counter so that the for statement executes once for each number of years. Calculating the pay raise. In the body of the for statement, display the value of counter in the first column and the value of wage in the second column. Then calculate the new weekly wage for the following year, and store the resulting value in the wage variable. To do this, add 1 to the percentage increase (be sure to divide the percentage by 100.0 ) and multiply the result by the current value in wage. Save, compile and run the application. Input a raise percentage and a number of years for the wage increase. View the results to ensure that the correct years are displayed and that the future wage results are correct. Close the Command Prompt window. We can not figure it out! Any help would be greatly appreciated, thanks!

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  • Help me alter this query to get the desired results - New*

    - by sandeepan
    Please dump these data first CREATE TABLE IF NOT EXISTS `all_tag_relations` ( `id_tag_rel` int(10) NOT NULL AUTO_INCREMENT, `id_tag` int(10) unsigned NOT NULL DEFAULT '0', `id_tutor` int(10) DEFAULT NULL, `id_wc` int(10) unsigned DEFAULT NULL, PRIMARY KEY (`id_tag_rel`), KEY `All_Tag_Relations_FKIndex1` (`id_tag`), KEY `id_wc` (`id_wc`) ) ENGINE=InnoDB DEFAULT CHARSET=latin1 AUTO_INCREMENT=19 ; INSERT INTO `all_tag_relations` (`id_tag_rel`, `id_tag`, `id_tutor`, `id_wc`) VALUES (1, 1, 1, NULL), (2, 2, 1, NULL), (3, 6, 2, NULL), (4, 7, 2, NULL), (8, 3, 1, 1), (9, 4, 1, 1), (10, 5, 2, 2), (11, 4, 2, 2), (15, 8, 1, 3), (16, 9, 1, 3), (17, 10, 1, 4), (18, 4, 1, 4), (19, 1, 2, 5), (20, 4, 2, 5); CREATE TABLE IF NOT EXISTS `tags` ( `id_tag` int(10) unsigned NOT NULL AUTO_INCREMENT, `tag` varchar(255) DEFAULT NULL, PRIMARY KEY (`id_tag`), UNIQUE KEY `tag` (`tag`), KEY `id_tag` (`id_tag`), KEY `tag_2` (`tag`), KEY `tag_3` (`tag`), KEY `tag_4` (`tag`), FULLTEXT KEY `tag_5` (`tag`) ) ENGINE=MyISAM DEFAULT CHARSET=latin1 AUTO_INCREMENT=11 ; INSERT INTO `tags` (`id_tag`, `tag`) VALUES (1, 'Sandeepan'), (2, 'Nath'), (3, 'first'), (4, 'class'), (5, 'new'), (6, 'Bob'), (7, 'Cratchit'), (8, 'more'), (9, 'fresh'), (10, 'second'); CREATE TABLE IF NOT EXISTS `webclasses` ( `id_wc` int(10) NOT NULL AUTO_INCREMENT, `id_author` int(10) NOT NULL, `name` varchar(50) DEFAULT NULL, PRIMARY KEY (`id_wc`) ) ENGINE=InnoDB DEFAULT CHARSET=latin1 AUTO_INCREMENT=5 ; INSERT INTO `webclasses` (`id_wc`, `id_author`, `name`) VALUES (1, 1, 'first class'), (2, 2, 'new class'), (3, 1, 'more fresh'), (4, 1, 'second class'), (5, 2, 'sandeepan class'); About the system - The system consists of tutors and classes. - The data in the table All_Tag_Relations stores tag relations for each tutor registered and each class created by a tutor. The tag relations are used for searching classes. The current data dump corresponds to tutor "Sandeepan Nath" who has created classes named "first class", "more fresh", "second class" and tutor "Bob Cratchit" who has created classes "new class" and "Sandeepan class". I am trying for a search query performs AND logic on the search keywords and returns wvery such class for which the search terms are present in the class name or its tutor name To make it easy, following is the list of search terms and desired results:- Search term result classes (check the id_wc in the results) first class 1 Sandeepan Nath class 1 Sandeepan Nath 1,3 Bob Cratchit 2 Sandeepan Nath bob none Sandeepan Class 1,4,5 I have so far reached upto this query -- Two keywords search SET @tag1 = 4, @tag2 = 1; -- Setting some user variables to see where the ids go. SELECT wc.id_wc, sum( DISTINCT ( wtagrels.id_tag = @tag1 ) ) AS key_1_class_matches, sum( DISTINCT ( wtagrels.id_tag = @tag2 ) ) AS key_2_class_matches, sum( DISTINCT ( ttagrels.id_tag = @tag1 ) ) AS key_1_tutor_matches, sum( DISTINCT ( ttagrels.id_tag = @tag2 ) ) AS key_2_tutor_matches, sum( DISTINCT ( ttagrels.id_tag = wtagrels.id_tag ) ) AS key_class_tutor_matches FROM WebClasses as wc join all_tag_relations AS wtagrels on wc.id_wc = wtagrels.id_wc join all_tag_relations as ttagrels on (wc.id_author = ttagrels.id_tutor) WHERE ( wtagrels.id_tag = @tag1 OR wtagrels.id_tag = @tag2 OR ttagrels.id_tag = @tag1 OR ttagrels.id_tag = @tag2 ) GROUP BY wtagrels.id_wc LIMIT 0 , 20 For search with 1 or 3 terms, remove/add the variable part in this query. Tabulating my observation of the values of key_1_class_matches, key_2_class_matches,key_1_tutor_matches (say, class keys),key_2_tutor_matches for various cases (say, tutor keys). Search term expected result Observation first class 1 for class 1, all class keys+all tutor keys =1 Sandeepan Nath class 1 for class 1, one class key+ all tutor keys = 1 Sandeepan Nath 1,3 both tutor keys =1 for these classes Bob Cratchit 2 both tutor keys = 1 Sandeepan Nath bob none no complete tutor matches for any class I found a pattern that, for any case, the class(es) which should appear in the result have the highest number of matches (all class keys and tutor keys). E.g. searching "first class", only for class =1, total of key matches = 4(1+1+1+1) searching "Sandeepan Nath", for classes 1, 3,4(all classes by Sandeepan Nath) have all the tutor keys matching. But no pattern in the search for "Sandeepan Class" - classes 1,4,5 should match. Now, how do I put a condition into the query, based on that pattern so that only those classes are returned. Do I need to use full text search here because it gives a scoring/rank value indicating the strength of the match? Any sample query would help. Please note - I have already found solution for showing classes when any/all of the search terms match with the class name. http://stackoverflow.com/questions/3030022/mysql-help-me-alter-this-search-query-to-get-desired-results But if all the search terms are in tutor name, it does not work. So, I am modifying the query and experimenting.

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