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

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

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  • Creating Custom Ajax Control Toolkit Controls

    - by Stephen Walther
    The goal of this blog entry is to explain how you can extend the Ajax Control Toolkit with custom Ajax Control Toolkit controls. I describe how you can create the two halves of an Ajax Control Toolkit control: the server-side control extender and the client-side control behavior. Finally, I explain how you can use the new Ajax Control Toolkit control in a Web Forms page. At the end of this blog entry, there is a link to download a Visual Studio 2010 solution which contains the code for two Ajax Control Toolkit controls: SampleExtender and PopupHelpExtender. The SampleExtender contains the minimum skeleton for creating a new Ajax Control Toolkit control. You can use the SampleExtender as a starting point for your custom Ajax Control Toolkit controls. The PopupHelpExtender control is a super simple custom Ajax Control Toolkit control. This control extender displays a help message when you start typing into a TextBox control. The animated GIF below demonstrates what happens when you click into a TextBox which has been extended with the PopupHelp extender. Here’s a sample of a Web Forms page which uses the control: <%@ Page Language="C#" AutoEventWireup="true" CodeBehind="ShowPopupHelp.aspx.cs" Inherits="MyACTControls.Web.Default" %> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html > <head runat="server"> <title>Show Popup Help</title> </head> <body> <form id="form1" runat="server"> <div> <act:ToolkitScriptManager ID="tsm" runat="server" /> <%-- Social Security Number --%> <asp:Label ID="lblSSN" Text="SSN:" AssociatedControlID="txtSSN" runat="server" /> <asp:TextBox ID="txtSSN" runat="server" /> <act:PopupHelpExtender id="ph1" TargetControlID="txtSSN" HelpText="Please enter your social security number." runat="server" /> <%-- Social Security Number --%> <asp:Label ID="lblPhone" Text="Phone Number:" AssociatedControlID="txtPhone" runat="server" /> <asp:TextBox ID="txtPhone" runat="server" /> <act:PopupHelpExtender id="ph2" TargetControlID="txtPhone" HelpText="Please enter your phone number." runat="server" /> </div> </form> </body> </html> In the page above, the PopupHelp extender is used to extend the functionality of the two TextBox controls. When focus is given to a TextBox control, the popup help message is displayed. An Ajax Control Toolkit control extender consists of two parts: a server-side control extender and a client-side behavior. For example, the PopupHelp extender consists of a server-side PopupHelpExtender control (PopupHelpExtender.cs) and a client-side PopupHelp behavior JavaScript script (PopupHelpBehavior.js). Over the course of this blog entry, I describe how you can create both the server-side extender and the client-side behavior. Writing the Server-Side Code Creating a Control Extender You create a control extender by creating a class that inherits from the abstract ExtenderControlBase class. For example, the PopupHelpExtender control is declared like this: public class PopupHelpExtender: ExtenderControlBase { } The ExtenderControlBase class is part of the Ajax Control Toolkit. This base class contains all of the common server properties and methods of every Ajax Control Toolkit extender control. The ExtenderControlBase class inherits from the ExtenderControl class. The ExtenderControl class is a standard class in the ASP.NET framework located in the System.Web.UI namespace. This class is responsible for generating a client-side behavior. The class generates a call to the Microsoft Ajax Library $create() method which looks like this: <script type="text/javascript"> $create(MyACTControls.PopupHelpBehavior, {"HelpText":"Please enter your social security number.","id":"ph1"}, null, null, $get("txtSSN")); }); </script> The JavaScript $create() method is part of the Microsoft Ajax Library. The reference for this method can be found here: http://msdn.microsoft.com/en-us/library/bb397487.aspx This method accepts the following parameters: type – The type of client behavior to create. The $create() method above creates a client PopupHelpBehavior. Properties – Enables you to pass initial values for the properties of the client behavior. For example, the initial value of the HelpText property. This is how server property values are passed to the client. Events – Enables you to pass client-side event handlers to the client behavior. References – Enables you to pass references to other client components. Element – The DOM element associated with the client behavior. This will be the DOM element associated with the control being extended such as the txtSSN TextBox. The $create() method is generated for you automatically. You just need to focus on writing the server-side control extender class. Specifying the Target Control All Ajax Control Toolkit extenders inherit a TargetControlID property from the ExtenderControlBase class. This property, the TargetControlID property, points at the control that the extender control extends. For example, the Ajax Control Toolkit TextBoxWatermark control extends a TextBox, the ConfirmButton control extends a Button, and the Calendar control extends a TextBox. You must indicate the type of control which your extender is extending. You indicate the type of control by adding a [TargetControlType] attribute to your control. For example, the PopupHelp extender is declared like this: [TargetControlType(typeof(TextBox))] public class PopupHelpExtender: ExtenderControlBase { } The PopupHelp extender can be used to extend a TextBox control. If you try to use the PopupHelp extender with another type of control then an exception is thrown. If you want to create an extender control which can be used with any type of ASP.NET control (Button, DataView, TextBox or whatever) then use the following attribute: [TargetControlType(typeof(Control))] Decorating Properties with Attributes If you decorate a server-side property with the [ExtenderControlProperty] attribute then the value of the property gets passed to the control’s client-side behavior. The value of the property gets passed to the client through the $create() method discussed above. The PopupHelp control contains the following HelpText property: [ExtenderControlProperty] [RequiredProperty] public string HelpText { get { return GetPropertyValue("HelpText", "Help Text"); } set { SetPropertyValue("HelpText", value); } } The HelpText property determines the help text which pops up when you start typing into a TextBox control. Because the HelpText property is decorated with the [ExtenderControlProperty] attribute, any value assigned to this property on the server is passed to the client automatically. For example, if you declare the PopupHelp extender in a Web Form page like this: <asp:TextBox ID="txtSSN" runat="server" /> <act:PopupHelpExtender id="ph1" TargetControlID="txtSSN" HelpText="Please enter your social security number." runat="server" />   Then the PopupHelpExtender renders the call to the the following Microsoft Ajax Library $create() method: $create(MyACTControls.PopupHelpBehavior, {"HelpText":"Please enter your social security number.","id":"ph1"}, null, null, $get("txtSSN")); You can see this call to the JavaScript $create() method by selecting View Source in your browser. This call to the $create() method calls a method named set_HelpText() automatically and passes the value “Please enter your social security number”. There are several attributes which you can use to decorate server-side properties including: ExtenderControlProperty – When a property is marked with this attribute, the value of the property is passed to the client automatically. ExtenderControlEvent – When a property is marked with this attribute, the property represents a client event handler. Required – When a value is not assigned to this property on the server, an error is displayed. DefaultValue – The default value of the property passed to the client. ClientPropertyName – The name of the corresponding property in the JavaScript behavior. For example, the server-side property is named ID (uppercase) and the client-side property is named id (lower-case). IDReferenceProperty – Applied to properties which refer to the IDs of other controls. URLProperty – Calls ResolveClientURL() to convert from a server-side URL to a URL which can be used on the client. ElementReference – Returns a reference to a DOM element by performing a client $get(). The WebResource, ClientResource, and the RequiredScript Attributes The PopupHelp extender uses three embedded resources named PopupHelpBehavior.js, PopupHelpBehavior.debug.js, and PopupHelpBehavior.css. The first two files are JavaScript files and the final file is a Cascading Style sheet file. These files are compiled as embedded resources. You don’t need to mark them as embedded resources in your Visual Studio solution because they get added to the assembly when the assembly is compiled by a build task. You can see that these files get embedded into the MyACTControls assembly by using Red Gate’s .NET Reflector tool: In order to use these files with the PopupHelp extender, you need to work with both the WebResource and the ClientScriptResource attributes. The PopupHelp extender includes the following three WebResource attributes. [assembly: WebResource("PopupHelp.PopupHelpBehavior.js", "text/javascript")] [assembly: WebResource("PopupHelp.PopupHelpBehavior.debug.js", "text/javascript")] [assembly: WebResource("PopupHelp.PopupHelpBehavior.css", "text/css", PerformSubstitution = true)] These WebResource attributes expose the embedded resource from the assembly so that they can be accessed by using the ScriptResource.axd or WebResource.axd handlers. The first parameter passed to the WebResource attribute is the name of the embedded resource and the second parameter is the content type of the embedded resource. The PopupHelp extender also includes the following ClientScriptResource and ClientCssResource attributes: [ClientScriptResource("MyACTControls.PopupHelpBehavior", "PopupHelp.PopupHelpBehavior.js")] [ClientCssResource("PopupHelp.PopupHelpBehavior.css")] Including these attributes causes the PopupHelp extender to request these resources when you add the PopupHelp extender to a page. If you open View Source in a browser which uses the PopupHelp extender then you will see the following link for the Cascading Style Sheet file: <link href="/WebResource.axd?d=0uONMsWXUuEDG-pbJHAC1kuKiIMteQFkYLmZdkgv7X54TObqYoqVzU4mxvaa4zpn5H9ch0RDwRYKwtO8zM5mKgO6C4WbrbkWWidKR07LD1d4n4i_uNB1mHEvXdZu2Ae5mDdVNDV53znnBojzCzwvSw2&amp;t=634417392021676003" type="text/css" rel="stylesheet" /> You also will see the following script include for the JavaScript file: <script src="/ScriptResource.axd?d=pIS7xcGaqvNLFBvExMBQSp_0xR3mpDfS0QVmmyu1aqDUjF06TrW1jVDyXNDMtBHxpRggLYDvgFTWOsrszflZEDqAcQCg-hDXjun7ON0Ol7EXPQIdOe1GLMceIDv3OeX658-tTq2LGdwXhC1-dE7_6g2&amp;t=ffffffff88a33b59" type="text/javascript"></script> The JavaScrpt file returned by this request to ScriptResource.axd contains the combined scripts for any and all Ajax Control Toolkit controls in a page. By default, the Ajax Control Toolkit combines all of the JavaScript files required by a page into a single JavaScript file. Combining files in this way really speeds up how quickly all of the JavaScript files get delivered from the web server to the browser. So, by default, there will be only one ScriptResource.axd include for all of the JavaScript files required by a page. If you want to disable Script Combining, and create separate links, then disable Script Combining like this: <act:ToolkitScriptManager ID="tsm" runat="server" CombineScripts="false" /> There is one more important attribute used by Ajax Control Toolkit extenders. The PopupHelp behavior uses the following two RequirdScript attributes to load the JavaScript files which are required by the PopupHelp behavior: [RequiredScript(typeof(CommonToolkitScripts), 0)] [RequiredScript(typeof(PopupExtender), 1)] The first parameter of the RequiredScript attribute represents either the string name of a JavaScript file or the type of an Ajax Control Toolkit control. The second parameter represents the order in which the JavaScript files are loaded (This second parameter is needed because .NET attributes are intrinsically unordered). In this case, the RequiredScript attribute will load the JavaScript files associated with the CommonToolkitScripts type and the JavaScript files associated with the PopupExtender in that order. The PopupHelp behavior depends on these JavaScript files. Writing the Client-Side Code The PopupHelp extender uses a client-side behavior written with the Microsoft Ajax Library. Here is the complete code for the client-side behavior: (function () { // The unique name of the script registered with the // client script loader var scriptName = "PopupHelpBehavior"; function execute() { Type.registerNamespace('MyACTControls'); MyACTControls.PopupHelpBehavior = function (element) { /// <summary> /// A behavior which displays popup help for a textbox /// </summmary> /// <param name="element" type="Sys.UI.DomElement">The element to attach to</param> MyACTControls.PopupHelpBehavior.initializeBase(this, [element]); this._textbox = Sys.Extended.UI.TextBoxWrapper.get_Wrapper(element); this._cssClass = "ajax__popupHelp"; this._popupBehavior = null; this._popupPosition = Sys.Extended.UI.PositioningMode.BottomLeft; this._popupDiv = null; this._helpText = "Help Text"; this._element$delegates = { focus: Function.createDelegate(this, this._element_onfocus), blur: Function.createDelegate(this, this._element_onblur) }; } MyACTControls.PopupHelpBehavior.prototype = { initialize: function () { MyACTControls.PopupHelpBehavior.callBaseMethod(this, 'initialize'); // Add event handlers for focus and blur var element = this.get_element(); $addHandlers(element, this._element$delegates); }, _ensurePopup: function () { if (!this._popupDiv) { var element = this.get_element(); var id = this.get_id(); this._popupDiv = $common.createElementFromTemplate({ nodeName: "div", properties: { id: id + "_popupDiv" }, cssClasses: ["ajax__popupHelp"] }, element.parentNode); this._popupBehavior = new $create(Sys.Extended.UI.PopupBehavior, { parentElement: element }, {}, {}, this._popupDiv); this._popupBehavior.set_positioningMode(this._popupPosition); } }, get_HelpText: function () { return this._helpText; }, set_HelpText: function (value) { if (this._HelpText != value) { this._helpText = value; this._ensurePopup(); this._popupDiv.innerHTML = value; this.raisePropertyChanged("Text") } }, _element_onfocus: function (e) { this.show(); }, _element_onblur: function (e) { this.hide(); }, show: function () { this._popupBehavior.show(); }, hide: function () { if (this._popupBehavior) { this._popupBehavior.hide(); } }, dispose: function() { var element = this.get_element(); $clearHandlers(element); if (this._popupBehavior) { this._popupBehavior.dispose(); this._popupBehavior = null; } } }; MyACTControls.PopupHelpBehavior.registerClass('MyACTControls.PopupHelpBehavior', Sys.Extended.UI.BehaviorBase); Sys.registerComponent(MyACTControls.PopupHelpBehavior, { name: "popupHelp" }); } // execute if (window.Sys && Sys.loader) { Sys.loader.registerScript(scriptName, ["ExtendedBase", "ExtendedCommon"], execute); } else { execute(); } })();   In the following sections, we’ll discuss how this client-side behavior works. Wrapping the Behavior for the Script Loader The behavior is wrapped with the following script: (function () { // The unique name of the script registered with the // client script loader var scriptName = "PopupHelpBehavior"; function execute() { // Behavior Content } // execute if (window.Sys && Sys.loader) { Sys.loader.registerScript(scriptName, ["ExtendedBase", "ExtendedCommon"], execute); } else { execute(); } })(); This code is required by the Microsoft Ajax Library Script Loader. You need this code if you plan to use a behavior directly from client-side code and you want to use the Script Loader. If you plan to only use your code in the context of the Ajax Control Toolkit then you can leave out this code. Registering a JavaScript Namespace The PopupHelp behavior is declared within a namespace named MyACTControls. In the code above, this namespace is created with the following registerNamespace() method: Type.registerNamespace('MyACTControls'); JavaScript does not have any built-in way of creating namespaces to prevent naming conflicts. The Microsoft Ajax Library extends JavaScript with support for namespaces. You can learn more about the registerNamespace() method here: http://msdn.microsoft.com/en-us/library/bb397723.aspx Creating the Behavior The actual Popup behavior is created with the following code. MyACTControls.PopupHelpBehavior = function (element) { /// <summary> /// A behavior which displays popup help for a textbox /// </summmary> /// <param name="element" type="Sys.UI.DomElement">The element to attach to</param> MyACTControls.PopupHelpBehavior.initializeBase(this, [element]); this._textbox = Sys.Extended.UI.TextBoxWrapper.get_Wrapper(element); this._cssClass = "ajax__popupHelp"; this._popupBehavior = null; this._popupPosition = Sys.Extended.UI.PositioningMode.BottomLeft; this._popupDiv = null; this._helpText = "Help Text"; this._element$delegates = { focus: Function.createDelegate(this, this._element_onfocus), blur: Function.createDelegate(this, this._element_onblur) }; } MyACTControls.PopupHelpBehavior.prototype = { initialize: function () { MyACTControls.PopupHelpBehavior.callBaseMethod(this, 'initialize'); // Add event handlers for focus and blur var element = this.get_element(); $addHandlers(element, this._element$delegates); }, _ensurePopup: function () { if (!this._popupDiv) { var element = this.get_element(); var id = this.get_id(); this._popupDiv = $common.createElementFromTemplate({ nodeName: "div", properties: { id: id + "_popupDiv" }, cssClasses: ["ajax__popupHelp"] }, element.parentNode); this._popupBehavior = new $create(Sys.Extended.UI.PopupBehavior, { parentElement: element }, {}, {}, this._popupDiv); this._popupBehavior.set_positioningMode(this._popupPosition); } }, get_HelpText: function () { return this._helpText; }, set_HelpText: function (value) { if (this._HelpText != value) { this._helpText = value; this._ensurePopup(); this._popupDiv.innerHTML = value; this.raisePropertyChanged("Text") } }, _element_onfocus: function (e) { this.show(); }, _element_onblur: function (e) { this.hide(); }, show: function () { this._popupBehavior.show(); }, hide: function () { if (this._popupBehavior) { this._popupBehavior.hide(); } }, dispose: function() { var element = this.get_element(); $clearHandlers(element); if (this._popupBehavior) { this._popupBehavior.dispose(); this._popupBehavior = null; } } }; The code above has two parts. The first part of the code is used to define the constructor function for the PopupHelp behavior. This is a factory method which returns an instance of a PopupHelp behavior: MyACTControls.PopupHelpBehavior = function (element) { } The second part of the code modified the prototype for the PopupHelp behavior: MyACTControls.PopupHelpBehavior.prototype = { } Any code which is particular to a single instance of the PopupHelp behavior should be placed in the constructor function. For example, the default value of the _helpText field is assigned in the constructor function: this._helpText = "Help Text"; Any code which is shared among all instances of the PopupHelp behavior should be added to the PopupHelp behavior’s prototype. For example, the public HelpText property is added to the prototype: get_HelpText: function () { return this._helpText; }, set_HelpText: function (value) { if (this._HelpText != value) { this._helpText = value; this._ensurePopup(); this._popupDiv.innerHTML = value; this.raisePropertyChanged("Text") } }, Registering a JavaScript Class After you create the PopupHelp behavior, you must register the behavior as a class by using the Microsoft Ajax registerClass() method like this: MyACTControls.PopupHelpBehavior.registerClass('MyACTControls.PopupHelpBehavior', Sys.Extended.UI.BehaviorBase); This call to registerClass() registers PopupHelp behavior as a class which derives from the base Sys.Extended.UI.BehaviorBase class. Like the ExtenderControlBase class on the server side, the BehaviorBase class on the client side contains method used by every behavior. The documentation for the BehaviorBase class can be found here: http://msdn.microsoft.com/en-us/library/bb311020.aspx The most important methods and properties of the BehaviorBase class are the following: dispose() – Use this method to clean up all resources used by your behavior. In the case of the PopupHelp behavior, the dispose() method is used to remote the event handlers created by the behavior and disposed the Popup behavior. get_element() -- Use this property to get the DOM element associated with the behavior. In other words, the DOM element which the behavior extends. get_id() – Use this property to the ID of the current behavior. initialize() – Use this method to initialize the behavior. This method is called after all of the properties are set by the $create() method. Creating Debug and Release Scripts You might have noticed that the PopupHelp behavior uses two scripts named PopupHelpBehavior.js and PopupHelpBehavior.debug.js. However, you never create these two scripts. Instead, you only create a single script named PopupHelpBehavior.pre.js. The pre in PopupHelpBehavior.pre.js stands for preprocessor. When you build the Ajax Control Toolkit (or the sample Visual Studio Solution at the end of this blog entry), a build task named JSBuild generates the PopupHelpBehavior.js release script and PopupHelpBehavior.debug.js debug script automatically. The JSBuild preprocessor supports the following directives: #IF #ELSE #ENDIF #INCLUDE #LOCALIZE #DEFINE #UNDEFINE The preprocessor directives are used to mark code which should only appear in the debug version of the script. The directives are used extensively in the Microsoft Ajax Library. For example, the Microsoft Ajax Library Array.contains() method is created like this: $type.contains = function Array$contains(array, item) { //#if DEBUG var e = Function._validateParams(arguments, [ {name: "array", type: Array, elementMayBeNull: true}, {name: "item", mayBeNull: true} ]); if (e) throw e; //#endif return (indexOf(array, item) >= 0); } Notice that you add each of the preprocessor directives inside a JavaScript comment. The comment prevents Visual Studio from getting confused with its Intellisense. The release version, but not the debug version, of the PopupHelpBehavior script is also minified automatically by the Microsoft Ajax Minifier. The minifier is invoked by a build step in the project file. Conclusion The goal of this blog entry was to explain how you can create custom AJAX Control Toolkit controls. In the first part of this blog entry, you learned how to create the server-side portion of an Ajax Control Toolkit control. You learned how to derive a new control from the ExtenderControlBase class and decorate its properties with the necessary attributes. Next, in the second part of this blog entry, you learned how to create the client-side portion of an Ajax Control Toolkit control by creating a client-side behavior with JavaScript. You learned how to use the methods of the Microsoft Ajax Library to extend your client behavior from the BehaviorBase class. Download the Custom ACT Starter Solution

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  • Much Ado About Nothing: Stub Objects

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

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  • Towards Database Continuous Delivery – What Next after Continuous Integration? A Checklist

    - by Ben Rees
    .dbd-banner p{ font-size:0.75em; padding:0 0 10px; margin:0 } .dbd-banner p span{ color:#675C6D; } .dbd-banner p:last-child{ padding:0; } @media ALL and (max-width:640px){ .dbd-banner{ background:#f0f0f0; padding:5px; color:#333; margin-top: 5px; } } -- Database delivery patterns & practices STAGE 4 AUTOMATED DEPLOYMENT If you’ve been fortunate enough to get to the stage where you’ve implemented some sort of continuous integration process for your database updates, then hopefully you’re seeing the benefits of that investment – constant feedback on changes your devs are making, advanced warning of data loss (prior to the production release on Saturday night!), a nice suite of automated tests to check business logic, so you know it’s going to work when it goes live, and so on. But what next? What can you do to improve your delivery process further, moving towards a full continuous delivery process for your database? In this article I describe some of the issues you might need to tackle on the next stage of this journey, and how to plan to overcome those obstacles before they appear. Our Database Delivery Learning Program consists of four stages, really three – source controlling a database, running continuous integration processes, then how to set up automated deployment (the middle stage is split in two – basic and advanced continuous integration, making four stages in total). If you’ve managed to work through the first three of these stages – source control, basic, then advanced CI, then you should have a solid change management process set up where, every time one of your team checks in a change to your database (whether schema or static reference data), this change gets fully tested automatically by your CI server. But this is only part of the story. Great, we know that our updates work, that the upgrade process works, that the upgrade isn’t going to wipe our 4Tb of production data with a single DROP TABLE. But – how do you get this (fully tested) release live? Continuous delivery means being always ready to release your software at any point in time. There’s a significant gap between your latest version being tested, and it being easily releasable. Just a quick note on terminology – there’s a nice piece here from Atlassian on the difference between continuous integration, continuous delivery and continuous deployment. This piece also gives a nice description of the benefits of continuous delivery. These benefits have been summed up by Jez Humble at Thoughtworks as: “Continuous delivery is a set of principles and practices to reduce the cost, time, and risk of delivering incremental changes to users” There’s another really useful piece here on Simple-Talk about the need for continuous delivery and how it applies to the database written by Phil Factor – specifically the extra needs and complexities of implementing a full CD solution for the database (compared to just implementing CD for, say, a web app). So, hopefully you’re convinced of moving on the the next stage! The next step after CI is to get some sort of automated deployment (or “release management”) process set up. But what should I do next? What do I need to plan and think about for getting my automated database deployment process set up? Can’t I just install one of the many release management tools available and hey presto, I’m ready! If only it were that simple. Below I list some of the areas that it’s worth spending a little time on, where a little planning and prep could go a long way. It’s also worth pointing out, that this should really be an evolving process. Depending on your starting point of course, it can be a long journey from your current setup to a full continuous delivery pipeline. If you’ve got a CI mechanism in place, you’re certainly a long way down that path. Nevertheless, we’d recommend evolving your process incrementally. Pages 157 and 129-141 of the book on Continuous Delivery (by Jez Humble and Dave Farley) have some great guidance on building up a pipeline incrementally: http://www.amazon.com/Continuous-Delivery-Deployment-Automation-Addison-Wesley/dp/0321601912 For now, in this post, we’ll look at the following areas for your checklist: You and Your Team Environments The Deployment Process Rollback and Recovery Development Practices You and Your Team It’s a cliché in the DevOps community that “It’s not all about processes and tools, really it’s all about a culture”. As stated in this DevOps report from Puppet Labs: “DevOps processes and tooling contribute to high performance, but these practices alone aren’t enough to achieve organizational success. The most common barriers to DevOps adoption are cultural: lack of manager or team buy-in, or the value of DevOps isn’t understood outside of a specific group”. Like most clichés, there’s truth in there – if you want to set up a database continuous delivery process, you need to get your boss, your department, your company (if relevant) onside. Why? Because it’s an investment with the benefits coming way down the line. But the benefits are huge – for HP, in the book A Practical Approach to Large-Scale Agile Development: How HP Transformed LaserJet FutureSmart Firmware, these are summarized as: -2008 to present: overall development costs reduced by 40% -Number of programs under development increased by 140% -Development costs per program down 78% -Firmware resources now driving innovation increased by a factor of 8 (from 5% working on new features to 40% But what does this mean? It means that, when moving to the next stage, to make that extra investment in automating your deployment process, it helps a lot if everyone is convinced that this is a good thing. That they understand the benefits of automated deployment and are willing to make the effort to transform to a new way of working. Incidentally, if you’re ever struggling to convince someone of the value I’d strongly recommend just buying them a copy of this book – a great read, and a very practical guide to how it can really work at a large org. I’ve spoken to many customers who have implemented database CI who describe their deployment process as “The point where automation breaks down. Up to that point, the CI process runs, untouched by human hand, but as soon as that’s finished we revert to manual.” This deployment process can involve, for example, a DBA manually comparing an environment (say, QA) to production, creating the upgrade scripts, reading through them, checking them against an Excel document emailed to him/her the night before, turning to page 29 in his/her notebook to double-check how replication is switched off and on for deployments, and so on and so on. Painful, error-prone and lengthy. But the point is, if this is something like your deployment process, telling your DBA “We’re changing everything you do and your toolset next week, to automate most of your role – that’s okay isn’t it?” isn’t likely to go down well. There’s some work here to bring him/her onside – to explain what you’re doing, why there will still be control of the deployment process and so on. Or of course, if you’re the DBA looking after this process, you have to do a similar job in reverse. You may have researched and worked out how you’d like to change your methodology to start automating your painful release process, but do the dev team know this? What if they have to start producing different artifacts for you? Will they be happy with this? Worth talking to them, to find out. As well as talking to your DBA/dev team, the other group to get involved before implementation is your manager. And possibly your manager’s manager too. As mentioned, unless there’s buy-in “from the top”, you’re going to hit problems when the implementation starts to get rocky (and what tool/process implementations don’t get rocky?!). You need to have support from someone senior in your organisation – someone you can turn to when you need help with a delayed implementation, lack of resources or lack of progress. Actions: Get your DBA involved (or whoever looks after live deployments) and discuss what you’re planning to do or, if you’re the DBA yourself, get the dev team up-to-speed with your plans, Get your boss involved too and make sure he/she is bought in to the investment. Environments Where are you going to deploy to? And really this question is – what environments do you want set up for your deployment pipeline? Assume everyone has “Production”, but do you have a QA environment? Dedicated development environments for each dev? Proper pre-production? I’ve seen every setup under the sun, and there is often a big difference between “What we want, to do continuous delivery properly” and “What we’re currently stuck with”. Some of these differences are: What we want What we’ve got Each developer with their own dedicated database environment A single shared “development” environment, used by everyone at once An Integration box used to test the integration of all check-ins via the CI process, along with a full suite of unit-tests running on that machine In fact if you have a CI process running, you’re likely to have some sort of integration server running (even if you don’t call it that!). Whether you have a full suite of unit tests running is a different question… Separate QA environment used explicitly for manual testing prior to release “We just test on the dev environments, or maybe pre-production” A proper pre-production (or “staging”) box that matches production as closely as possible Hopefully a pre-production box of some sort. But does it match production closely!? A production environment reproducible from source control A production box which has drifted significantly from anything in source control The big question is – how much time and effort are you going to invest in fixing these issues? In reality this just involves figuring out which new databases you’re going to create and where they’ll be hosted – VMs? Cloud-based? What about size/data issues – what data are you going to include on dev environments? Does it need to be masked to protect access to production data? And often the amount of work here really depends on whether you’re working on a new, greenfield project, or trying to update an existing, brownfield application. There’s a world if difference between starting from scratch with 4 or 5 clean environments (reproducible from source control of course!), and trying to re-purpose and tweak a set of existing databases, with all of their surrounding processes and quirks. But for a proper release management process, ideally you have: Dedicated development databases, An Integration server used for testing continuous integration and running unit tests. [NB: This is the point at which deployments are automatic, without human intervention. Each deployment after this point is a one-click (but human) action], QA – QA engineers use a one-click deployment process to automatically* deploy chosen releases to QA for testing, Pre-production. The environment you use to test the production release process, Production. * A note on the use of the word “automatic” – when carrying out automated deployments this does not mean that the deployment is happening without human intervention (i.e. that something is just deploying over and over again). It means that the process of carrying out the deployment is automatic in that it’s not a person manually running through a checklist or set of actions. The deployment still requires a single-click from a user. Actions: Get your environments set up and ready, Set access permissions appropriately, Make sure everyone understands what the environments will be used for (it’s not a “free-for-all” with all environments to be accessed, played with and changed by development). The Deployment Process As described earlier, most existing database deployment processes are pretty manual. The following is a description of a process we hear very often when we ask customers “How do your database changes get live? How does your manual process work?” Check pre-production matches production (use a schema compare tool, like SQL Compare). Sometimes done by taking a backup from production and restoring in to pre-prod, Again, use a schema compare tool to find the differences between the latest version of the database ready to go live (i.e. what the team have been developing). This generates a script, User (generally, the DBA), reviews the script. This often involves manually checking updates against a spreadsheet or similar, Run the script on pre-production, and check there are no errors (i.e. it upgrades pre-production to what you hoped), If all working, run the script on production.* * this assumes there’s no problem with production drifting away from pre-production in the interim time period (i.e. someone has hacked something in to the production box without going through the proper change management process). This difference could undermine the validity of your pre-production deployment test. Red Gate is currently working on a free tool to detect this problem – sign up here at www.sqllighthouse.com, if you’re interested in testing early versions. There are several variations on this process – some better, some much worse! How do you automate this? In particular, step 3 – surely you can’t automate a DBA checking through a script, that everything is in order!? The key point here is to plan what you want in your new deployment process. There are so many options. At one extreme, pure continuous deployment – whenever a dev checks something in to source control, the CI process runs (including extensive and thorough testing!), before the deployment process keys in and automatically deploys that change to the live box. Not for the faint hearted – and really not something we recommend. At the other extreme, you might be more comfortable with a semi-automated process – the pre-production/production matching process is automated (with an error thrown if these environments don’t match), followed by a manual intervention, allowing for script approval by the DBA. One he/she clicks “Okay, I’m happy for that to go live”, the latter stages automatically take the script through to live. And anything in between of course – and other variations. But we’d strongly recommended sitting down with a whiteboard and your team, and spending a couple of hours mapping out “What do we do now?”, “What do we actually want?”, “What will satisfy our needs for continuous delivery, but still maintaining some sort of continuous control over the process?” NB: Most of what we’re discussing here is about production deployments. It’s important to note that you will also need to map out a deployment process for earlier environments (for example QA). However, these are likely to be less onerous, and many customers opt for a much more automated process for these boxes. Actions: Sit down with your team and a whiteboard, and draw out the answers to the questions above for your production deployments – “What do we do now?”, “What do we actually want?”, “What will satisfy our needs for continuous delivery, but still maintaining some sort of continuous control over the process?” Repeat for earlier environments (QA and so on). Rollback and Recovery If only every deployment went according to plan! Unfortunately they don’t – and when things go wrong, you need a rollback or recovery plan for what you’re going to do in that situation. Once you move in to a more automated database deployment process, you’re far more likely to be deploying more frequently than before. No longer once every 6 months, maybe now once per week, or even daily. Hence the need for a quick rollback or recovery process becomes paramount, and should be planned for. NB: These are mainly scenarios for handling rollbacks after the transaction has been committed. If a failure is detected during the transaction, the whole transaction can just be rolled back, no problem. There are various options, which we’ll explore in subsequent articles, things like: Immediately restore from backup, Have a pre-tested rollback script (remembering that really this is a “roll-forward” script – there’s not really such a thing as a rollback script for a database!) Have fallback environments – for example, using a blue-green deployment pattern. Different options have pros and cons – some are easier to set up, some require more investment in infrastructure; and of course some work better than others (the key issue with using backups, is loss of the interim transaction data that has been added between the failed deployment and the restore). The best mechanism will be primarily dependent on how your application works and how much you need a cast-iron failsafe mechanism. Actions: Work out an appropriate rollback strategy based on how your application and business works, your appetite for investment and requirements for a completely failsafe process. Development Practices This is perhaps the more difficult area for people to tackle. The process by which you can deploy database updates is actually intrinsically linked with the patterns and practices used to develop that database and linked application. So you need to decide whether you want to implement some changes to the way your developers actually develop the database (particularly schema changes) to make the deployment process easier. A good example is the pattern “Branch by abstraction”. Explained nicely here, by Martin Fowler, this is a process that can be used to make significant database changes (e.g. splitting a table) in a step-wise manner so that you can always roll back, without data loss – by making incremental updates to the database backward compatible. Slides 103-108 of the following slidedeck, from Niek Bartholomeus explain the process: https://speakerdeck.com/niekbartho/orchestration-in-meatspace As these slides show, by making a significant schema change in multiple steps – where each step can be rolled back without any loss of new data – this affords the release team the opportunity to have zero-downtime deployments with considerably less stress (because if an increment goes wrong, they can roll back easily). There are plenty more great patterns that can be implemented – the book Refactoring Databases, by Scott Ambler and Pramod Sadalage is a great read, if this is a direction you want to go in: http://www.amazon.com/Refactoring-Databases-Evolutionary-paperback-Addison-Wesley/dp/0321774515 But the question is – how much of this investment are you willing to make? How often are you making significant schema changes that would require these best practices? Again, there’s a difference here between migrating old projects and starting afresh – with the latter it’s much easier to instigate best practice from the start. Actions: For your business, work out how far down the path you want to go, amending your database development patterns to “best practice”. It’s a trade-off between implementing quality processes, and the necessity to do so (depending on how often you make complex changes). Socialise these changes with your development group. No-one likes having “best practice” changes imposed on them, so good to introduce these ideas and the rationale behind them early.   Summary The next stages of implementing a continuous delivery pipeline for your database changes (once you have CI up and running) require a little pre-planning, if you want to get the most out of the work, and for the implementation to go smoothly. We’ve covered some of the checklist of areas to consider – mainly in the areas of “Getting the team ready for the changes that are coming” and “Planning our your pipeline, environments, patterns and practices for development”, though there will be more detail, depending on where you’re coming from – and where you want to get to. This article is part of our database delivery patterns & practices series on Simple Talk. Find more articles for version control, automated testing, continuous integration & deployment.

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  • apache eats up too much ram per child

    - by mrc4r7m4n
    Hello to everyone. I've got fallowing problem: Apache eat to many ram per child. The fallowing comments shows: cat /etc/redhat-release -- Fedora release 8 (Werewolf) free -m: total used free shared buffers cached Mem: 3566 3136 429 0 339 1907 -/+ buffers/cache: 889 2676 Swap: 4322 0 4322 I know that you will say that there is nothing to worry about because swap is not use, but i think it's not use for now. 3.httpd -v: Server version: Apache/2.2.14 (Unix) 4.httpd -l: Compiled in modules: core.c mod_authn_file.c mod_authn_default.c mod_authz_host.c mod_authz_groupfile.c mod_authz_user.c mod_authz_default.c mod_auth_basic.c mod_include.c mod_filter.c mod_log_config.c mod_env.c mod_setenvif.c mod_version.c mod_ssl.c prefork.c http_core.c mod_mime.c mod_status.c mod_autoindex.c mod_asis.c mod_cgi.c mod_negotiation.c mod_dir.c mod_actions.c mod_userdir.c mod_alias.c mod_rewrite.c mod_so.c 5.List of loaded dynamic modules: LoadModule authz_host_module modules/mod_authz_host.so LoadModule include_module modules/mod_include.so LoadModule log_config_module modules/mod_log_config.so LoadModule setenvif_module modules/mod_setenvif.so LoadModule mime_module modules/mod_mime.so LoadModule autoindex_module modules/mod_autoindex.so LoadModule vhost_alias_module modules/mod_vhost_alias.so LoadModule negotiation_module modules/mod_negotiation.so LoadModule dir_module modules/mod_dir.so LoadModule alias_module modules/mod_alias.so LoadModule rewrite_module modules/mod_rewrite.so LoadModule proxy_module modules/mod_proxy.so LoadModule cgi_module modules/mod_cgi.so 6.My prefrok directive <IfModule prefork.c> StartServers 8 MinSpareServers 5 MaxSpareServers 25 ServerLimit 80 MaxClients 80 MaxRequestsPerChild 4000 </IfModule> KeepAliveTimeout 6 MaxKeepAliveRequests 100 KeepAlive On 7.top -u apache: ctrl+ M top - 09:19:42 up 2 days, 19 min, 2 users, load average: 0.85, 0.87, 0.80 Tasks: 113 total, 1 running, 112 sleeping, 0 stopped, 0 zombie Cpu(s): 7.3%us, 15.7%sy, 0.0%ni, 75.7%id, 0.0%wa, 0.7%hi, 0.7%si, 0.0%st Mem: 3652120k total, 3149964k used, 502156k free, 348048k buffers Swap: 4425896k total, 0k used, 4425896k free, 1944952k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 16956 apache 20 0 700m 135m 100m S 0.0 3.8 2:16.78 httpd 16953 apache 20 0 565m 130m 96m S 0.0 3.7 1:57.26 httpd 16957 apache 20 0 587m 129m 102m S 0.0 3.6 1:47.41 httpd 16955 apache 20 0 567m 126m 93m S 0.0 3.6 1:43.60 httpd 17494 apache 20 0 626m 125m 96m S 0.0 3.5 1:58.77 httpd 17515 apache 20 0 540m 120m 88m S 0.0 3.4 1:45.57 httpd 17516 apache 20 0 573m 120m 88m S 0.0 3.4 1:50.51 httpd 16954 apache 20 0 551m 120m 88m S 0.0 3.4 1:52.47 httpd 17493 apache 20 0 586m 120m 94m S 0.0 3.4 1:51.02 httpd 17279 apache 20 0 568m 117m 87m S 16.0 3.3 1:51.87 httpd 17302 apache 20 0 560m 116m 90m S 0.3 3.3 1:59.06 httpd 17495 apache 20 0 551m 116m 89m S 0.0 3.3 1:47.51 httpd 17277 apache 20 0 476m 114m 81m S 0.0 3.2 1:37.14 httpd 30097 apache 20 0 536m 113m 83m S 0.0 3.2 1:47.38 httpd 30112 apache 20 0 530m 112m 81m S 0.0 3.2 1:40.15 httpd 17513 apache 20 0 516m 112m 85m S 0.0 3.1 1:43.92 httpd 16958 apache 20 0 554m 111m 82m S 0.0 3.1 1:44.18 httpd 1617 apache 20 0 487m 111m 85m S 0.0 3.1 1:31.67 httpd 16952 apache 20 0 461m 107m 75m S 0.0 3.0 1:13.71 httpd 16951 apache 20 0 462m 103m 76m S 0.0 2.9 1:28.05 httpd 17278 apache 20 0 497m 103m 76m S 0.0 2.9 1:31.25 httpd 17403 apache 20 0 537m 102m 79m S 0.0 2.9 1:52.24 httpd 25081 apache 20 0 412m 101m 70m S 0.0 2.8 1:01.74 httpd I guess thats all information needed to help me solve this problem. I think the virt memory is to big, the same res. The consumption of ram is increasing all the time. Maybe it's memory leak because i see there is so many static modules compiled. Could someone help me with this issue? Thank you in advance. 8.ldd /usr/sbin/httpd linux-gate.so.1 => (0x0012d000) libm.so.6 => /lib/libm.so.6 (0x0012e000) libpcre.so.0 => /lib/libpcre.so.0 (0x00157000) libselinux.so.1 => /lib/libselinux.so.1 (0x0017f000) libaprutil-1.so.0 => /usr/lib/libaprutil-1.so.0 (0x0019a000) libcrypt.so.1 => /lib/libcrypt.so.1 (0x001b4000) libldap-2.3.so.0 => /usr/lib/libldap-2.3.so.0 (0x001e6000) liblber-2.3.so.0 => /usr/lib/liblber-2.3.so.0 (0x00220000) libdb-4.6.so => /lib/libdb-4.6.so (0x0022e000) libexpat.so.1 => /lib/libexpat.so.1 (0x00370000) libapr-1.so.0 => /usr/lib/libapr-1.so.0 (0x00391000) libpthread.so.0 => /lib/libpthread.so.0 (0x003b9000) libdl.so.2 => /lib/libdl.so.2 (0x003d2000) libc.so.6 => /lib/libc.so.6 (0x003d7000) /lib/ld-linux.so.2 (0x00110000) libuuid.so.1 => /lib/libuuid.so.1 (0x00530000) libresolv.so.2 => /lib/libresolv.so.2 (0x00534000) libsasl2.so.2 => /usr/lib/libsasl2.so.2 (0x00548000) libssl.so.6 => /lib/libssl.so.6 (0x00561000) libcrypto.so.6 => /lib/libcrypto.so.6 (0x005a6000) libgssapi_krb5.so.2 => /usr/lib/libgssapi_krb5.so.2 (0x006d9000) libkrb5.so.3 => /usr/lib/libkrb5.so.3 (0x00707000) libcom_err.so.2 => /lib/libcom_err.so.2 (0x0079a000) libk5crypto.so.3 => /usr/lib/libk5crypto.so.3 (0x0079d000) libz.so.1 => /lib/libz.so.1 (0x007c3000) libkrb5support.so.0 => /usr/lib/libkrb5support.so.0 (0x007d6000) libkeyutils.so.1 => /lib/libkeyutils.so.1 (0x007df000) Currently i cant restart the apache. I work in a company and now there is rush hours. I will do that about 5 pm. Current top -u apache: shift + M top - 12:31:33 up 2 days, 3:30, 1 user, load average: 0.73, 0.80, 0.79 Tasks: 114 total, 1 running, 113 sleeping, 0 stopped, 0 zombie Cpu(s): 3.3%us, 4.7%sy, 0.0%ni, 90.0%id, 1.3%wa, 0.3%hi, 0.3%si, 0.0%st Mem: 3652120k total, 3169720k used, 482400k free, 353372k buffers Swap: 4425896k total, 0k used, 4425896k free, 1978688k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 16957 apache 20 0 708m 145m 117m S 0.0 4.1 2:11.32 httpd 16956 apache 20 0 754m 142m 107m S 0.0 4.0 2:33.94 httpd 16955 apache 20 0 641m 136m 103m S 5.3 3.8 1:58.37 httpd 17515 apache 20 0 624m 131m 99m S 0.0 3.7 2:03.90 httpd 16954 apache 20 0 627m 130m 98m S 0.0 3.6 2:13.87 httpd 17302 apache 20 0 625m 124m 97m S 0.0 3.5 2:10.80 httpd 17403 apache 20 0 624m 114m 91m S 0.0 3.2 2:08.85 httpd 16952 apache 20 0 502m 114m 81m S 0.0 3.2 1:23.78 httpd 16186 apache 20 0 138m 61m 35m S 0.0 1.7 0:15.54 httpd 16169 apache 20 0 111m 49m 17m S 0.0 1.4 0:06.00 httpd 16190 apache 20 0 126m 48m 24m S 0.0 1.4 0:11.44 httpd 16191 apache 20 0 109m 48m 19m S 0.0 1.4 0:04.62 httpd 16163 apache 20 0 114m 48m 21m S 0.0 1.4 0:09.60 httpd 16183 apache 20 0 127m 48m 23m S 0.0 1.3 0:11.23 httpd 16189 apache 20 0 109m 47m 17m S 0.0 1.3 0:04.55 httpd 16201 apache 20 0 106m 47m 17m S 0.0 1.3 0:03.90 httpd 16193 apache 20 0 103m 46m 20m S 0.0 1.3 0:10.76 httpd 16188 apache 20 0 107m 45m 18m S 0.0 1.3 0:04.85 httpd 16168 apache 20 0 103m 44m 17m S 0.0 1.2 0:05.61 httpd 16187 apache 20 0 118m 41m 21m S 0.0 1.2 0:08.50 httpd 16184 apache 20 0 111m 41m 19m S 0.0 1.2 0:09.28 httpd 16206 apache 20 0 110m 41m 20m S 0.0 1.2 0:11.69 httpd 16199 apache 20 0 108m 40m 17m S 0.0 1.1 0:07.76 httpd 16166 apache 20 0 104m 37m 18m S 0.0 1.0 0:04.31 httpd 16185 apache 20 0 99.3m 36m 16m S 0.0 1.0 0:04.16 httpd as you can see the memory usage growing up from e.g. res( 135 to 145)m and it will be growing up till memory ends. Are you sure that this option i set up: <IfModule prefork.c> StartServers 8 MinSpareServers 5 MaxSpareServers 25 ServerLimit 80 MaxClients 80 MaxRequestsPerChild 4000 </IfModule> KeepAliveTimeout 6 MaxKeepAliveRequests 100 KeepAlive On are correct? Maybe i should decrease some of them? Another questions that bother me: I got e.g. static module mod_negotiation.c compiled into apache and the same module loaded as dynamic. Is this normal that i've loaded duplicated module. But when i want to remove dynamic module(mod_negotiation.c) from httpd.conf and then restart apache error appears. Now I cant tell this error message because i cant restart apache :( Hello again:) This is memory usage just after restart apache: top - 16:19:12 up 2 days, 7:18, 3 users, load average: 1.08, 0.91, 0.91 Tasks: 109 total, 2 running, 107 sleeping, 0 stopped, 0 zombie Cpu(s): 17.0%us, 25.7%sy, 51.0%ni, 4.7%id, 0.0%wa, 0.3%hi, 1.3%si, 0.0%st Mem: 3652120k total, 2762516k used, 889604k free, 361552k buffers Swap: 4425896k total, 0k used, 4425896k free, 2020980k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 13569 apache 20 0 93416 43m 15m S 0.0 1.2 0:02.55 httpd 13575 apache 20 0 98356 38m 16m S 32.3 1.1 0:02.55 httpd 13571 apache 20 0 86808 33m 12m S 0.0 0.9 0:02.60 httpd 13568 apache 20 0 86760 33m 12m S 0.0 0.9 0:00.81 httpd 13570 apache 20 0 83480 33m 12m S 0.0 0.9 0:00.51 httpd 13572 apache 20 0 63520 5916 1548 S 0.0 0.2 0:00.02 httpd 13573 apache 20 0 63520 5916 1548 S 0.0 0.2 0:00.02 httpd 13574 apache 20 0 63520 5916 1548 S 0.0 0.2 0:00.02 httpd 13761 apache 20 0 63388 5128 860 S 0.0 0.1 0:00.01 httpd 13762 apache 20 0 63388 5128 860 S 0.0 0.1 0:00.01 httpd 13763 apache 20 0 63388 5128 860 S 0.0 0.1 0:00.00 httpd I will try to compile apache from source to newest version. Thx for help guys.

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  • Windows 7 Samba issue

    - by abduls85
    We have a strange samba issue affecting only one user. Our samba setup is as follow : Red Hat Enterprise Linux Server release 5.4 (Tikanga) - Samba Server Samba version 3.0.33-3.14.el5 - Samba version Domain Controller WIN2008R2 Standard - Windows DC Windows 7 64 bit - Client PCs User mentioned that he faced this problem after he force shutdown his PC few weeks ago. By right, for all users when we access \\sambaservername in windows it will show all the shares in the samba server but for this user once he startup his PC he will not be able to access \\sambaservername, Error message Windows cannot access \\sambaservername Current workaround to solve the problem : Try to access one share in \\sambaservername for instance \\sambaservername\sharedfolder1. But even when doing so, it will first prompt an error in the beginning, error message is as follows Logon failure: unknown user name or bad password. user need to enter the credentials again and he can access the share. Thereafter, he will be able to access \\sambaservername without any issues. But once he reboots his computer the problem will persists. Troubleshooting done so far: Ensure the following settings: Go to: Control Panel → Administrative Tools → Local Security Policy Select: Local Policies → Security Options "Network security: LAN Manager authentication level" → Send LM & NTLM responses "Minimum session security for NTLM SSP" → uncheck: Require 128-bit encryption Advise user to reset his password and try again but problem still persists Tried my account on users' PC, there is no issues. Tried user account on serveral other Windows 7 PC including mine but problem still persists. Windows XP does not have this problem. Ensure that there is no stored crendentials on the windows 7 PC. Checked the credentials manager in Control Panel as well as typing this command rundll32.exe keymgr.dll, KRShowKeyMgr Restart winbindd daemon on samba server but to no avail. I suspect this is due to some caching issue but not sure where is the issue. Whenever the user has error accessing \\sambaservername, the following errors will be logged in the samba server : [2012/10/10 17:10:26, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! [2012/10/10 17:10:27, 1] smbd/sesssetup.c:reply_spnego_kerberos(316) Failed to verify incoming ticket with error NT_STATUS_LOGON_FAILURE! But after workaround, there will be no more errors. I suspect after reading the article listed below some amendments need to be made to the \var\samba\cache directory : http://www.linuxquestions.org/questions/linux-server-73/getent-passwd-dont-show-ad-groups-and-users-745829/ http://www.samba.org/samba/docs/man/Samba-HOWTO-Collection/tdb.html http://lists.samba.org/archive/samba/2010-May/155521.html http://lists.samba.org/archive/samba/2011-March/161912.html http://lzeit.blogspot.sg/2009/10/samba-shares-inaccessible-after-power.html There are several users using the samba server and i would like to solve this problem without any impacts. I saw the following article : http://www.samba.org/samba/docs/man/manpages-3/smb.conf.5.html#WINBINDCACHETIME "winbind offline logon (G) This parameter is designed to control whether Winbind should allow to login with the pam_winbind module using Cached Credentials. If enabled, winbindd will store user credentials from successful logins encrypted in a local cache. Default: winbind offline logon = false Example: winbind offline logon = true " Any idea on how to delete the entry for one user in the local cache ?

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  • Why can't I build Deluge?

    - by hugemeow
    Deluge is a BitTorrent Client. I am trying to build it from source, since I don't have privilege to install it as root. I am using python setup.py build. But, it failed following message, why? copying deluge/ui/web/themes/images/gray/slider/slider-v-thumb.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/gray/slider copying deluge/ui/web/themes/images/gray/slider/slider-thumb.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/gray/slider copying deluge/ui/web/themes/images/gray/panel/top-bottom.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/gray/panel copying deluge/ui/web/themes/images/gray/tabs/tab-strip-bg.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/gray/tabs copying deluge/ui/web/themes/images/yourtheme/window/right-corners.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/window copying deluge/ui/web/themes/images/yourtheme/window/left-corners.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/window copying deluge/ui/web/themes/images/yourtheme/window/left-right.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/window copying deluge/ui/web/themes/images/yourtheme/window/top-bottom.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/window creating build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/slider copying deluge/ui/web/themes/images/yourtheme/slider/slider-v-thumb.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/slider copying deluge/ui/web/themes/images/yourtheme/slider/slider-thumb.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/slider copying deluge/ui/web/themes/images/yourtheme/slider/slider-bg.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/slider copying deluge/ui/web/themes/images/yourtheme/slider/slider-v-bg.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/slider copying deluge/ui/web/themes/images/yourtheme/panel/top-bottom.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/panel copying deluge/ui/web/themes/images/yourtheme/grid/hmenu-lock.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/grid copying deluge/ui/web/themes/images/yourtheme/grid/hmenu-unlock.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/grid copying deluge/ui/web/themes/images/yourtheme/tabs/tab-strip-bg.png -> build/lib.linux-x86_64-2.4/deluge/ui/web/themes/images/yourtheme/tabs running build_ext building 'libtorrent' extension gcc -pthread -shared -L/usr/lib64 -L/opt/local/lib -lboost_filesystem -lboost_date_time -lboost_iostreams -lboost_python -lboost_thread -lpthread -lssl -lz -o build/lib.linux-x86_64-2.4/deluge/libtorrent.so /usr/bin/ld: cannot find -lboost_filesystem collect2: ld returned 1 exit status error: command 'gcc' failed with exit status 1 [mirror@innov deluge-1.3.5]$ echo $? 1 Edit 1: gcc version and os information $(which gcc) --version gcc (GCC) 4.1.2 20080704 (Red Hat 4.1.2-52) Copyright (C) 2006 Free Software Foundation, Inc. This is free software; see the source for copying conditions. There is NO warranty; not even for MERCHANTABILITY or FITNESS FOR A PARTICULAR PURPOSE. cat /etc/issue CentOS release 5.7 (Final) Kernel \r on an \m Edit 2: boost is referenced by setup.py in deluge 114 if OS == "linux": 115 if os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 116 'libboost_filesystem-mt.so')): 117 boost_filesystem = "boost_filesystem-mt" 118 elif os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 119 'libboost_filesystem.so')): 120 boost_filesystem = "boost_filesystem" 121 if os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 122 'libboost_date_time-mt.so')): 123 boost_date_time = "boost_date_time-mt" 124 elif os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 125 'libboost_date_time.so')): 126 boost_date_time = "boost_date_time" 127 if os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 128 'libboost_thread-mt.so')): 129 boost_thread = "boost_thread-mt" 130 elif os.path.exists(os.path.join(sysconfig.get_config_vars()['LIBDIR'], \ 131 'libboost_thread.so')): 132 boost_thread = "boost_thread" 133 134 if 'boost_filesystem' not in vars(): 135 boost_filesystem = "boost_filesystem-mt" 136 if 'boost_date_time' not in vars(): 137 boost_date_time = "boost_date_time-mt" 138 if 'boost_thread' not in vars(): 139 boost_thread = "boost_thread-mt" 140 141 elif OS == "freebsd": 142 boost_filesystem = "boost_filesystem" 143 boost_date_time = "boost_date_time" 144 boost_thread = "boost_thread" 145 else: 146 boost_filesystem = "boost_filesystem-mt" 147 boost_date_time = "boost_date_time-mt" 148 boost_thread = "boost_thread-mt" 149 150 librariestype = [boost_filesystem, boost_date_time, 151 boost_thread, 'z', 'pthread', 'ssl', 'crypto']

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  • How do I achieve lossless JPEG joining without truncation of partial MCUs?

    - by Karan
    I am working on a project for which I need to join thousands of JPEG images losslessly (I'm not talking about the Lossless JPEG/JPEG 2000/JPEG-LS formats here). Aforementioned images have varying levels of chroma subsampling (1x1, 1x2, 2x1, 2x2), resulting in varying MCU sizes (8x8, 8x16, 16x8, 16x16 px). However, in any given set of images to be joined together, each image has identical characteristics. For now, let's assume I only have 2 images. Image #1 (I1) is 256x256px in size and #2 (I2) is 239x256px in size. 2x2 subsampling is used such that MCU size is 16x16px. I2 thus obviously has partial MCUs at the right edge, since its width is not evenly divisible by 16. (I've read that so-called 'partial' MCUs actually contain the data for a complete MCU, but the image dimensions instruct the renderer to only display the relevant pixels and ignore/hide the extra ones.) Looking around for tools that could help me accomplish this, I came across a modified version of JpegTran, that contains an experimental lossless crop 'n' drop (cut & paste) feature. All the other apps I encountered that support lossless JPEG editing seem to utilise IJG's (JpegTran) code, so this seemed to be the logical choice. Also, given the sheer number of images, I wanted something that could preferably be run from the command-line so that I could automate the process with a script. Unfortunately, while everything else worked fine, it seems JpegTran truncates the partial MCUs instead of retaining them. Thus in the example above, the final joined image contains all of I1, but only 224x256px of I2. Why 224? because 239 = 14x16+15, which means there are 14 full MCUs along the width, and 1 partial MCU (just 1px short of the complete 16px). The last 15px is what is getting blanked, leading to a 495x256px image with 15px of blank (grey) pixels at the right edge. See images below (shame that imgur re-compresses them): (left )+ (right) = As you can clearly see, the red portion (15px) of I2 has been truncated by JpegTran. If the MCUs were 8px in width, the lost portion would have been the right-most 7px of I2. Similarly, joining I3 (256x239px) *below * I1 would cause the loss of 7 or 15px, depending on the MCU height of course: (top) + (bottom) = If this is better suited to some other StackExchange (or even non-SE) site/forum where JPEG/image encoding experts hang out, do let me know. Can what I am attempting even be done, or is the so-called 'lossless' JPEG crop 'n' drop only valid for images with no partial MCUs? (Maybe that is why the feature is still in an "experimental state" more than a decade after being introduced...) Until I know for sure that it is impossible, I am not interested in suggestions for lossy joining. Avoiding any generational loss whatsoever is the sole reason why I'm breaking my head over this, else I'd have had this done and dusted ages ago. Also, I am not interested in suggestions related to switching image formats. I do not control the source of the images. If it can be done, how? Please keep in mind that any alternate apps suggested must ideally be capable of automation, given the requirements stated above. (But given how it's unlikely I'm even going to receive a useful answer given the constraints, I would be happy with any app suggestion just as long as it actually works. I can always look into an AutoIT/AHK script or something later to automate it.) I understand that an odd-sized final image might cause issues, so I am fully prepared to accept any solution, even if it results in blank (preferably black) padding pixels to the right/bottom. What I mean is, I don't care if I1 + I2 is 496x256px (1px padding) or even 512x256px (17px padding) in size, as long as the final image contains all the actual image data from both source images, and the entire process is lossless. Obviously the lesser the padding (if any), the better, but at this point any solution will do. A Windows-based solution would be perfect, but a Linux-based one would be entirely acceptable.

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  • Nagios: NRPE: Unable to read output, Can't find the reason, can you?

    - by Itai Ganot
    I have a Nagios server and a monitored server. On the monitored server: [root@Monitored ~]# netstat -an |grep :5666 tcp 0 0 0.0.0.0:5666 0.0.0.0:* LISTEN [root@Monitored ~]# locate check_kvm /usr/lib64/nagios/plugins/check_kvm [root@Monitored ~]# /usr/lib64/nagios/plugins/check_kvm -H localhost hosts:3 OK:3 WARN:0 CRIT:0 - ab2c7:running alpweb5:running istaweb5:running [root@Monitored ~]# /usr/lib64/nagios/plugins/check_nrpe -H localhost -c check_kvm NRPE: Unable to read output [root@Monitored ~]# /usr/lib64/nagios/plugins/check_nrpe -H localhost NRPE v2.14 [root@Monitored ~]# ps -ef |grep nrpe nagios 21178 1 0 16:11 ? 00:00:00 /usr/sbin/nrpe -c /etc/nagios/nrpe.cfg -d [root@Monitored ~]# On the Nagios server: [root@Nagios ~]# /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.159 -c check_kvm NRPE: Unable to read output [root@Nagios ~]# /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.159 NRPE v2.14 [root@Nagios ~]# When I check another server in the network using the same command it works: [root@Nagios ~]# /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.80 -c check_kvm hosts:4 OK:4 WARN:0 CRIT:0 - karmisoft:running ab2c4:running kidumim1:running travel2gether1:running [root@Nagios ~]# Running the check locally using Nagios account: [root@Monitored ~]# su - nagios -bash-4.1$ /usr/lib64/nagios/plugins/check_kvm hosts:3 OK:3 WARN:0 CRIT:0 - ab2c7:running alpweb5:running istaweb5:running -bash-4.1$ Running the check remotely from the Nagios server using Nagios account: -bash-4.1$ /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.159 -c check_kvm NRPE: Unable to read output -bash-4.1$ /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.159 NRPE v2.14 -bash-4.1$ Running the same check_kvm against a different server in the network using Nagios account: -bash-4.1$ /usr/lib64/nagios/plugins/check_nrpe -H 1.1.1.80 -c check_kvm hosts:4 OK:4 WARN:0 CRIT:0 - karmisoft:running ab2c4:running kidumim1:running travel2gether1:running -bash-4.1$ Permissions: -rwxr-xr-x. 1 root root 4684 2013-10-14 17:14 nrpe.cfg (aka /etc/nagios/nrpe.cfg) drwxrwxr-x. 3 nagios nagios 4096 2013-10-15 03:38 plugins (aka /usr/lib64/nagios/plugins) /etc/sudoers: [root@Monitored ~]# grep -i requiretty /etc/sudoers #Defaults requiretty iptables/selinux: [root@Monitored xinetd.d]# service iptables status iptables: Firewall is not running. [root@Monitored xinetd.d]# service ip6tables status ip6tables: Firewall is not running. [root@Monitored xinetd.d]# grep disable /etc/selinux/config # disabled - No SELinux policy is loaded. SELINUX=disabled [root@Monitored xinetd.d]# The command in /etc/nagios/nrpe.cfg is: [root@Monitored ~]# grep kvm /etc/nagios/nrpe.cfg command[check_kvm]=sudo /usr/lib64/nagios/plugins/check_kvm and the nagios user is added on /etc/sudoers: nagios ALL=(ALL) NOPASSWD:/usr/lib64/nagios/plugins/check_kvm nagios ALL=(ALL) NOPASSWD:/usr/lib64/nagios/plugins/check_nrpe The check_kvm is a shell script, looks like that: #!/bin/sh LIST=$(virsh list --all | sed '1,2d' | sed '/^$/d'| awk '{print $2":"$3}') if [ ! "$LIST" ]; then EXITVAL=3 #Status 3 = UNKNOWN (orange) echo "Unknown guests" exit $EXITVAL fi OK=0 WARN=0 CRIT=0 NUM=0 for host in $(echo $LIST) do name=$(echo $host | awk -F: '{print $1}') state=$(echo $host | awk -F: '{print $2}') NUM=$(expr $NUM + 1) case "$state" in running|blocked) OK=$(expr $OK + 1) ;; paused) WARN=$(expr $WARN + 1) ;; shutdown|shut*|crashed) CRIT=$(expr $CRIT + 1) ;; *) CRIT=$(expr $CRIT + 1) ;; esac done if [ "$NUM" -eq "$OK" ]; then EXITVAL=0 #Status 0 = OK (green) fi if [ "$WARN" -gt 0 ]; then EXITVAL=1 #Status 1 = WARNING (yellow) fi if [ "$CRIT" -gt 0 ]; then EXITVAL=2 #Status 2 = CRITICAL (red) fi echo hosts:$NUM OK:$OK WARN:$WARN CRIT:$CRIT - $LIST exit $EXITVAL Edit (10/22/13): Following all that, I am now able to get some response from the script: [root@Monitored ~]# /usr/lib64/nagios/plugins/check_nrpe -H localhost -c check_kvm Unknown guests [root@Monitored ~]# /usr/lib64/nagios/plugins/check_nrpe -H localhost NRPE v2.14 [root@Monitored ~]# /usr/lib64/nagios/plugins/check_kvm hosts:3 OK:3 WARN:0 CRIT:0 - ab2c7:running alpweb5:running istaweb5:running [root@Monitored ~]# su - nagios -bash-4.1$ /usr/lib64/nagios/plugins/check_kvm hosts:3 OK:3 WARN:0 CRIT:0 - ab2c7:running alpweb5:running istaweb5:running -bash-4.1$ /usr/lib64/nagios/plugins/check_nrpe -H localhost -c check_kvm Unknown guests -bash-4.1$ /usr/lib64/nagios/plugins/check_nrpe -H localhost NRPE v2.14 It seems like the problem is some how related to the check_nrpe command or something which is related to the nrpe installation on the server.

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  • DNS request timed out. timeout was 2 seconds

    - by sahil007
    i had setup bind dns server on centos. from local lan it will work fine but from remote when i tried to nslookup ..it will give reply like "DNS request timed out...timeout was 2 seconds." what is the problem? this is my bind config---- // Red Hat BIND Configuration Tool options { directory "/var/named"; dump-file "/var/named/data/cache_dump.db"; statistics-file "/var/named/data/named_stats.txt"; query-source address * port 53; }; controls { inet 127.0.0.1 allow {localhost; } keys {rndckey; }; }; acl internals { 127.0.0.0/8; 192.168.0.0/24; 10.0.0.0/8; }; view "internal" { match-clients { internals; }; recursion yes; zone "mydomain.com" { type master; file "mydomain.com.zone"; }; zone "0.168.192.in-addr.arpa" { type master; file "0.168.192.in-addr.arpa.zone"; }; zone "." IN { type hint; file "named.root"; }; zone "localdomain." IN { type master; file "localdomain.zone"; allow-update { none; }; }; zone "localhost." IN { type master; file "localhost.zone"; allow-update { none; }; }; zone "0.0.127.in-addr.arpa." IN { type master; file "named.local"; allow-update { none; }; }; zone "0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.0.ip6.arpa." I N { type master; file "named.ip6.local"; allow-update { none; }; }; zone "255.in-addr.arpa." IN { type master; file "named.broadcast"; allow-update { none; }; }; zone "0.in-addr.arpa." IN { type master; file "named.zero"; allow-update { none; }; }; }; view "external" { match-clients { any; }; recursion no; zone "mydomain.com" { type master; file "mydomain.com.zone"; // file "/var/named/chroot/var/named/mydomain.com.zone"; }; zone "0.168.192.in-addr.arpa" { type master; file "0.168.192.in-addr.arpa.zone"; }; }; include "/etc/rndc.key";

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  • How to configure DNS Server on Fedora

    - by user863873
    I want to learn how to configure my home PC server into a web server with domain and host. My IP is 109.99.141.133 and now points to a phpinfo page host on my home server. My registed domain is: anunta-anunturi.ro I searched for a tutorial and I've read that I have to configure /etc/named.conf and the file sources for the new zone that I create. So, from the tutorials, my /etc/named.conf looks like this: // // named.conf // // Provided by Red Hat bind package to configure the ISC BIND named(8) DNS // server as a caching only nameserver (as a localhost DNS resolver only). // // See /usr/share/doc/bind*/sample/ for example named configuration files. // options { listen-on port 53 { 127.0.0.1; }; listen-on-v6 port 53 { ::1; }; directory "/var/named"; dump-file "/var/named/data/cache_dump.db"; statistics-file "/var/named/data/named_stats.txt"; memstatistics-file "/var/named/data/named_mem_stats.txt"; allow-query { localhost; }; recursion yes; dnssec-enable yes; dnssec-validation yes; dnssec-lookaside auto; /* Path to ISC DLV key */ bindkeys-file "/etc/named.iscdlv.key"; managed-keys-directory "/var/named/dynamic"; }; logging { channel default_debug { file "data/named.run"; severity dynamic; }; }; zone "anunta-anunturi.ro" IN { type master; file "/etc/anunta-anunturi.db"; }; zone "." IN { type hint; file "named.ca"; }; include "/etc/named.rfc1912.zones"; include "/etc/named.root.key"; My /etc/anunta-anunturi.db file looks like this — I'm not sure if this is okay, or if it's the easy one. $TTL 86400 anunta-anunturi.ro. IN SOA serveur.anunta-anunturi.ro. root.serveur.anunta-anunturi.ro. ( 1997022700 ; Serial 28800 ; Refresh 14400 ; Retry 3600000 ; Expire 86400 ) ; Minumun IN NS serveur.anunta-anunturi.ro. IN MX 10 mail.anunta-anunturi.ro. serveur.anunta-anunturi.ro. IN A 192.168.1.37 www.anunta-anunturi.ro. IN A 192.168.1.37 mail.anunta-anunturi.ro. IN A 192.168.1.37 Extra info: At home I receive internet from my ISP through a router. My home PC and server recieve their IP automatically from the router when I start/restart. In my local home network, my server receives the IP 192.168.1.37 from the router. When I enter 109.99.141.133 in my browser, it points to the rooter that forwards port 80 to local IP 192.168.1.37 (my home server) Questions: Are my two files good? What/where is my nameserver that I need to copy/paste to my top level domain (where I registered my domain: rotld.ro)?

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  • PPTP connection fails with errors 800/806

    - by Mark S. Rasmussen
    I've got a client (Server 2008 R2) that won't connect to our production environment PPTP VPN server (Server 2003, running RRAS). The server is behind a firewall that has TCP1723 open as well as GRE. Other clients at our office are able to connect just fine. Our office is behind a Juniper SSG5-Serial firewall, but all outgoing traffic is allowed, and multiple other clients are able to connect to VPN servers without issues. I've also setup a completely different VPN server on another network outside of our office. The functioning clients connect just fine - the Server 2008 R2 machine doesn't. Thus it's definitely a problem with this machine in particular. I've rebooted it. I've disabled the firewall, no dice on either. I've run PPTPSRV and PPTPCLNT on the server/client and they're able to communicate perfectly - indicating there's no problem using neither TCP1723 nor GRE. The Server 2008 R2 machine is also running as a VPN server itself (incoming connection) and that's working perfectly. We have the issues no matter if there are active incoming connections or not. I'm not sure what my next debugging step would be; any suggestions? EDIT: The event log on the server has the following warning from RasMan: A connection between the VPN server and the VPN client xxx.xxx.xxx.xxx has been established, but the VPN connection cannot be completed. The most common cause for this is that a firewall or router between the VPN server and the VPN client is not configured to allow Generic Routing Encapsulation (GRE) packets (protocol 47). Verify that the firewalls and routers between your VPN server and the Internet allow GRE packets. Make sure the firewalls and routers on the user's network are also configured to allow GRE packets. If the problem persists, have the user contact the Internet service provider (ISP) to determine whether the ISP might be blocking GRE packets. Obviously this points to GRE being a potential problem. But seeing as I have other clients connectiong without problems, as well as PPTPSRV and PPTPCLNT being able to communicate, I'm suspecting this might be a red herring. EDIT: Here are the anonymized events logged by the client in chronological order: CoId={742CB15C-A7E0-47B7-8240-0EFA1139CBD9}: The user XXX\YYY has started dialing a VPN connection using a per-user connection profile named ZZZ. The connection settings are: Dial-in User = XXX\YYY VpnStrategy = PPTP DataEncryption = Require PrerequisiteEntry = AutoLogon = No UseRasCredentials = Yes Authentication Type = CHAP/MS-CHAPv2 Ipv4DefaultGateway = No Ipv4AddressAssignment = By Server Ipv4DNSServerAssignment = By Server Ipv6DefaultGateway = Yes Ipv6AddressAssignment = By Server Ipv6DNSServerAssignment = By Server IpDnsFlags = Register primary domain suffix IpNBTEnabled = Yes UseFlags = Private Connection ConnectOnWinlogon = No. CoId={742CB15C-A7E0-47B7-8240-0EFA1139CBD9}: The user XXX\YYY is trying to establish a link to the Remote Access Server for the connection named ZZZ using the following device: Server address/Phone Number = XXX.YYY.ZZZ.KKK Device = WAN Miniport (PPTP) Port = VPN3-4 MediaType = VPN. CoId={742CB15C-A7E0-47B7-8240-0EFA1139CBD9}: The user XXX\YYY has successfully established a link to the Remote Access Server using the following device: Server address/Phone Number = XXX.YYY.ZZZ.KKK Device = WAN Miniport (PPTP) Port = VPN3-4 MediaType = VPN. CoId={742CB15C-A7E0-47B7-8240-0EFA1139CBD9}: The link to the Remote Access Server has been established by user XXX\YYY. CoId={742CB15C-A7E0-47B7-8240-0EFA1139CBD9}: The user XXX\YYY dialed a connection named ZZZ which has failed. The error code returned on failure is 806. Running Wireshark on the client shows it trying and retrying to send a "71 Configuration Request" While the server shows the incoming client requests, but apparently without replying: Given that this is GRE traffic, I think rules out the GRE traffic being blocked. Question is, why doesn't the server reply? This is the Configuration Request the server receives from the non functioning client (meaning no response is sent to the client request): And this is the Configuration Request the server receives from the working client: To me they seem identical, except for differing keys and magic numbers, and the fact that one client receives a response while the other doesn't.

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  • Excel: conditionally format a cell using the format of another, content-matching cell

    - by Eric A. Meyer
    I have an Excel spreadsheet where I’d like to be able to create a “key” of formatted cells with unique values, and then in another sheet format cells using the key formatting. So for example, my key is as follows, with one value per cell and the visual formatting indicated in parentheses: A (red background) B (green background) C (blue background) So that’s on one sheet (or in a remote corner of the current sheet—whichever is better). Then, in an area that I mark for conditional formatting, I can type one of those three letters and have the cell where I typed it visually formatted according to the key. So if I type a “B” into one of the conditionally formatted cells, it gets a green background. (Note that I’m using backgrounds here solely for ease of explanation: ideally I want to have all visual formatting copied over, whether it’s foreground color, background color, font weight, borders, or whatever. But I’ll take what I can get, obviously.) And—just to make it extra-tricky—if I change the formatting in the key, that change should be reflected in cells that reference the key. Thus, if I change the “B” formatting in the key from a green background to a purple background, any “B” in the main sheet should switch to the new color. Similarly, it should be possible to add or remove values from the key and have those changes applied to the main data set. I’m okay with the formatting-update-on-key-change being triggered by clicking a button or something. I suspect that if any of this is possible it will require VBA, but I’ve never used it so I’ve no idea where to start if that’s the case. I’m hoping it’s possible without VBA. I know it’s possible to just use multiple conditional formats, but my use case here is that I’m trying to create the above-described capability for someone who isn’t conversant with conditional formatting. I’d like to let them be able to define a key, update it if necessary, and keep on truckin’ without me having to rewrite the spreadsheet’s formatting rules for them. --- UPDATE --- So I think I was a bit unclear about my original request. Let me try again with an image. The image shows the “key” on the left, where values and styles are defined using keyboard and mouse input. On the right, you see the data that should be formatted to match the key. Thus if I type a “C” into a cell in the Data area, it should be blue-backed. Furthermore, if I change the formatting of “C” in the Key to have a purple background, all the “C” cells should switch from blue to purple. For further craziness, if I add more to the Key (say, “D” with a yellow background) then any “D” cells will be styled to match; if I remove a Key entry, then matching values in the Data area should revert to default styling. So. Is that more clear? Is it possible, in whole or in part? I don’t have to use conditional formatting for this; in fact, at this point I suspect I probably shouldn’t. But I’m open to any approach!

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  • Crash in audio resampler with some audio rates - FFMPEG PHP ( Solved! )

    - by Olaf Erlandsen
    i have a problem with this command( FFMPEG PHP ): Command: ffmpeg -i 62f76f050494f0ed6a5997967c00c0c0.wmv -ss 0 -t 99 -y -ar 44100 -async 44100 -r 29.970 -ac 2 -qscale 5 -f flv 62f76f050494f0ed6a5997967c00c0c0.flv Output: FFmpeg version 0.6.5, Copyright (c) 2000-2010 the FFmpeg developers built on Jan 29 2012 17:52:15 with gcc 4.4.5 20110214 (Red Hat 4.4.5-6) configuration: --prefix=/usr --libdir=/usr/lib64 --shlibdir=/usr/lib64 --mandir=/usr/share/man --incdir=/usr/include --disable-avisynth --extra-cflags='-O2 -g -pipe -Wall -Wp,-D_FORTIFY_SOURCE=2 -fexceptions -fstack-protector --param=ssp-buffer-size=4 -m64 -mtune=generic -fPIC' --enable-avfilter --enable-avfilter-lavf --enable-libdc1394 --enable-libdirac --enable-libfaac --enable-libfaad --enable-libfaadbin --enable-libgsm --enable-libmp3lame --enable-libopencore-amrnb --enable-libopencore-amrwb --enable-librtmp --enable-libschroedinger --enable-libspeex --enable-libtheora --enable-libx264 --enable-gpl --enable-nonfree --enable-postproc --enable-pthreads --enable-shared --enable-swscale --enable-vdpau --enable-version3 --enable-x11grab libavutil 50.15. 1 / 50.15. 1 libavcodec 52.72. 2 / 52.72. 2 libavformat 52.64. 2 / 52.64. 2 libavdevice 52. 2. 0 / 52. 2. 0 libavfilter 1.19. 0 / 1.19. 0 libswscale 0.11. 0 / 0.11. 0 libpostproc 51. 2. 0 / 51. 2. 0 [asf @ 0xe81670]max_analyze_duration reached Input #0, asf, from '/var/www/resources/tmp/62f76f050494f0ed6a5997967c00c0c0.wmv': Metadata: WMFSDKVersion : 12.0.7601.17514 WMFSDKNeeded : 0.0.0.0000 IsVBR : 0 Duration: 00:00:50.87, bitrate: 2467 kb/s Stream #0.0: Audio: wmapro, 44100 Hz, stereo, flt, 256 kb/s Stream #0.1: Video: vc1, yuv420p, 950x460 [PAR 1:1 DAR 95:46], 25 fps, 25 tbr, 1k tbn, 25 tbc Output #0, flv, to '/var/www/resources/media/62f76f050494f0ed6a5997967c00c0c0.flv': Metadata: encoder : Lavf52.64.2 Stream #0.0: Video: flv, yuv420p, 950x460 [PAR 1:1 DAR 95:46], q=2-31, 200 kb/s, 1k tbn, 29.97 tbc Stream #0.1: Audio: libmp3lame, 11025 Hz, stereo, s16, 64 kb/s Stream mapping: Stream #0.1 -> #0.0 Stream #0.0 -> #0.1 Press [q] to stop encoding frame= 72 fps= 0 q=5.0 size= 0kB time=10.91 bitrate= 0.0kbits/s Multiple frames in a packet from stream 0 Warning, using s16 intermediate sample format for resampling frame= 141 fps=139 q=5.0 size= 103kB time=8.15 bitrate= 103.2kbits/s frame= 220 fps=144 q=5.0 size= 875kB time=10.92 bitrate= 656.6kbits/s frame= 290 fps=143 q=5.0 size= 1525kB time=13.74 bitrate= 909.1kbits/s frame= 356 fps=141 q=5.0 size= 2153kB time=15.99 bitrate=1103.1kbits/s frame= 427 fps=141 q=5.0 size= 2847kB time=18.70 bitrate=1247.0kbits/s frame= 497 fps=141 q=5.0 size= 3771kB time=21.16 bitrate=1460.0kbits/s frame= 575 fps=142 q=5.0 size= 4695kB time=24.61 bitrate=1563.0kbits/s frame= 639 fps=141 q=5.0 size= 5301kB time=26.80 bitrate=1620.2kbits/s frame= 703 fps=139 q=5.0 size= 5829kB time=29.36 bitrate=1626.2kbits/s frame= 774 fps=139 q=5.0 size= 6659kB time=32.39 bitrate=1684.0kbits/s frame= 842 fps=139 q=5.0 size= 7915kB time=35.27 bitrate=1838.6kbits/s frame= 911 fps=139 q=5.0 size= 9011kB time=37.98 bitrate=1943.4kbits/s frame= 975 fps=138 q=5.0 size= 9788kB time=40.59 bitrate=1975.3kbits/s frame= 1041 fps=138 q=5.0 size= 10904kB time=43.83 bitrate=2037.9kbits/s frame= 1115 fps=138 q=5.0 size= 11795kB time=46.24 bitrate=2089.8kbits/s frame= 1183 fps=138 q=5.0 size= 12678kB time=48.74 bitrate=2130.7kbits/s frame= 1247 fps=137 q=5.0 size= 13964kB time=51.36 bitrate=2227.5kbits/s frame= 1271 fps=136 q=5.0 Lsize= 15865kB time=58.86 bitrate=2208.1kbits/s video:15366kB audio:462kB global headers:0kB muxing overhead 0.238956% Problem: Warning, using s16 intermediate sample format for resampling I've also tried changing the parameter From -ar 44100 to -ar 11025 Thanks! Solution: Read this link: http://en.wikipedia.org/wiki/MP3#Bit_rate

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  • Unable to Mange DNS via MMC

    - by IT Helpdesk Team Manager
    When trying to access the DNS service on Microsoft Windows Server 2003 (Build 3790) domain controller/schema master via the MMC DNS snap in or locally via the DNS MMC from Administrative tools I'm getting a red "X" through the icon for the DNS Server. The inability to access DNS management via MMC happens on all domain controllers as well. We've looked at items such as the DHCP client not being started, incorrect DNS setup ( the machine points at itself and another DC ), the DNS service not running ( it is and all DNS queries via NSLOOKUP work correctly ), dslint returns the correct information and functions as expected. There is the following entry in the DNS event log: The DNS server could not initialize the remote procedure call (RPC) service. If it is not running, start the RPC service or reboot the computer. The event data is the error code. For more information, see Help and Support Center at http://go.microsoft.com/fwlink/events.asp. 0000: 0000051b dnscmd fails with RPC server unavailable yet RPC is started: C:\Documents and Settings\Administrator.DOMAIN>dnscmd /Info Info query failed status = 1722 (0x000006ba) Command failed: RPC_S_SERVER_UNAVAILABLE 1722 (000006ba) DCDIAG /TEST:DNS /V /E produces the following errors: Warning: no DNS RPC connectivity (error or non Microsoft DNS server is running) [Error details: 1753 (Type: Win32 - Description: There are no more endpoints available from the endpoint mapper.)] Warning: no DNS RPC connectivity (error or non Microsoft DNS server is running) [Error details: 1722 (Type: Win32 - Description: The RPC server is unavailable.)] The DNS server could not initialize the remote procedure call (RPC) service. If it is not running, start the RPC service or reboot the computer. The event data is the error code. A DNS query for _ldap._tcp.dc._msdcs. returns the correct results. All domain and ADS related activities are working except that I can't manage my DNS via MMC or dnscmd. Any thoughts or solutions would be greatly appreciated. EDIT: Adding Registry export per request: Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc Class Name: <NO CLASS> Last Write Time: 10/18/2012 - 2:29 PM Value 0 Name: DCOM Protocols Type: REG_MULTI_SZ Data: ncacn_ip_tcp Value 1 Name: UuidSequenceNumber Type: REG_DWORD Data: 0xb19bd0f Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc\ClientProtocols Class Name: <NO CLASS> Last Write Time: 3/9/2007 - 12:11 PM Value 0 Name: ncacn_np Type: REG_SZ Data: rpcrt4.dll Value 1 Name: ncacn_ip_tcp Type: REG_SZ Data: rpcrt4.dll Value 2 Name: ncadg_ip_udp Type: REG_SZ Data: rpcrt4.dll Value 3 Name: ncacn_http Type: REG_SZ Data: rpcrt4.dll Value 4 Name: ncacn_at_dsp Type: REG_SZ Data: rpcrt4.dll Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc\NameService Class Name: <NO CLASS> Last Write Time: 2/20/2006 - 4:48 PM Value 0 Name: DefaultSyntax Type: REG_SZ Data: 3 Value 1 Name: Endpoint Type: REG_SZ Data: \pipe\locator Value 2 Name: NetworkAddress Type: REG_SZ Data: \\. Value 3 Name: Protocol Type: REG_SZ Data: ncacn_np Value 4 Name: ServerNetworkAddress Type: REG_SZ Data: \\. Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc\NetBios Class Name: <NO CLASS> Last Write Time: 2/20/2006 - 4:48 PM Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc\RpcProxy Class Name: <NO CLASS> Last Write Time: 3/9/2007 - 12:11 PM Value 0 Name: Enabled Type: REG_DWORD Data: 0x1 Value 1 Name: ValidPorts Type: REG_SZ Data: pdc:100-5000 Key Name: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Rpc\SecurityService Class Name: <NO CLASS> Last Write Time: 2/20/2006 - 4:48 PM Value 0 Name: 9 Type: REG_SZ Data: secur32.dll Value 1 Name: 10 Type: REG_SZ Data: secur32.dll Value 2 Name: 14 Type: REG_SZ Data: schannel.dll Value 3 Name: 16 Type: REG_SZ Data: secur32.dll Value 4 Name: 1 Type: REG_SZ Data: secur32.dll Value 5 Name: 18 Type: REG_SZ Data: secur32.dll Value 6 Name: 68 Type: REG_SZ Data: netlogon.dll

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  • Incorrect gzipping of http requests, can't find who's doing it

    - by Ned Batchelder
    We're seeing some very strange mangling of HTTP responses, and we can't figure out what is doing it. We have an app server handling JSON requests. Occasionally, the response is returned gzipped, but with incorrect headers that prevent the browser from interpreting it correctly. The problem is intermittent, and changes behavior over time. Yesterday morning it seemed to fail 50% of the time, and in fact, seemed tied to one of our two load-balanced servers. Later in the afternoon, it was failing only 20 times out of 1000, and didn't correlate with an app server. The two app servers are running Apache 2.2 with mod_wsgi and a Django app stack. They have identical Apache configs and source trees, and even identical packages installed on Red Hat. There's a hardware load balancer in front, I don't know the make or model. Akamai is also part of the food chain, though we removed Akamai and still had the problem. Here's a good request and response: * Connected to example.com (97.7.79.129) port 80 (#0) > POST /claim/ HTTP/1.1 > User-Agent: curl/7.19.7 (x86_64-pc-linux-gnu) libcurl/7.19.7 OpenSSL/0.9.8k zlib/1.2.3.3 libidn/1.15 > Host: example.com > Accept: */* > Referer: http://example.com/apps/ > Accept-Encoding: gzip,deflate > Content-Length: 29 > Content-Type: application/x-www-form-urlencoded > } [data not shown] < HTTP/1.1 200 OK < Server: Apache/2 < Content-Language: en-us < Content-Encoding: identity < Content-Length: 47 < Content-Type: application/x-javascript < Connection: keep-alive < Vary: Accept-Encoding < { [data not shown] * Connection #0 to host example.com left intact * Closing connection #0 {"msg": "", "status": "OK", "printer_name": ""} And here's a bad one: * Connected to example.com (97.7.79.129) port 80 (#0) > POST /claim/ HTTP/1.1 > User-Agent: curl/7.19.7 (x86_64-pc-linux-gnu) libcurl/7.19.7 OpenSSL/0.9.8k zlib/1.2.3.3 libidn/1.15 > Host: example.com > Accept: */* > Referer: http://example.com/apps/ > Accept-Encoding: gzip,deflate > Content-Length: 29 > Content-Type: application/x-www-form-urlencoded > } [data not shown] < HTTP/1.1 200 OK < Server: Apache/2 < Content-Language: en-us < Content-Encoding: identity < Content-Type: application/x-javascript < Content-Encoding: gzip < Content-Length: 59 < Connection: keep-alive < Vary: Accept-Encoding < X-N: S < { [data not shown] * Connection #0 to host example.com left intact * Closing connection #0 ?V?-NW?RPR?QP*.I,)-???A??????????T??Z? ??/ There are two things to notice about the bad response: It has two Content-Encoding headers, and the browsers seem to use the first. So they see an identity encoding header, and gzipped content, so they can't interpret the response. The bad response has an extra "X-N: S" header. Perhaps if I could find out what intermediary adds "X-N: S" headers to responses, I could track down the culprit...

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  • Strange enduser experience with Liferay, Glassfish and Apache on RedHat

    - by Pete Helgren
    Tried multiple forums to get to the bottom of this. I hope I can get some direction here: Here is the stack I am working with: Red Hat Enterprise Linux Server release 5.6 (Tikanga) Liferay 6.0.6 on Glassfish 3.0.1 MySQL 5.0.77 Apache 2.2.3 The Liferay portal provides a variety of portlets to end users. Static content (web pages), static resources (primarily pdf and mp3 files 1mb - 80mb in size), File upload and download capabilities (primarily 40-60mb mp3 files) and online streaming of those MP3 files. Here is the strange end user experiences: Under normal load: (20-30) users uploading, downloading or streaming files and 20-30 accessing static content (some of it downloads), we see the following: 1) Clicking a link triggers the download of a portion of an MP3 (the portion is a few seconds long). 2) Clicking on a link tiggers the download of the page content rather than rendering. 3) Clicking a link causes the page to dump binary data to the end user rather than the expected content. 4) Clicking a link returns the text of a javascript file rather than rendering the page. Each occurrence is totally random (or appears so). Sometimes it works, sometimes it doesn't. It seems to have no relation to browser or client OS. The strange events seem to occur much more frequently when using an SSL connection rather than regular http. Apache serves as a proxy server only (reverse). It basically passes all the requests through to Glassfish. There isn't any static content proxy served by Apache. We rebuilt the entire stack from scratch and redeployed the portlet wars and still have the same issues. Liferay is running as a single server (not clustered). We disabled mod_cache in Apache. The problems are more frequent as the server load grows. This morning the load is pretty light and we are seeing few problems but the use of the site will grow, particularly tonight around 9pm CST through Wednesday morning. You could try the site (http://preview.bsfinternational.org) during those times and I would expect that you might experience one of the weirdnesses as you randomly click links on the site (https is invoked only when signed in). Again, https seems to exacerbate the issue. This seems very much like a caching issue but I don't know where in the stack to start peeling the onion. Apache? Liferay? Glassfish? MySQL? Maybe even Redhat? We are stumped and most forums we have posted to (LifeRay and Glassfish) have returned very few suggestions. I just need an idea of where to start looking. I understand that we could have a portlet EDIT: Opening the files in a Hex editor that appear to be pages that download rather than render, we see that the first 4000 characters are "junk" and then the "HTTP/1.1 ...." 'normal' header is seen. So something is dumping a jumble of characters up to offset 4000 (when viewing it in a Hex editor). Perhaps a clue? Ideas?

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  • Cannot connect to MySQL Server on RHEL 5.7

    - by Jeffrey Wong
    I have a standard MySQL Server running on Red hat 5.7. I have edited /etc/my.cnf to specify the bind address as my server's public IP address. [mysqld] datadir=/var/lib/mysql socket=/var/lib/mysql/mysql.sock user=mysql # Default to using old password format for compatibility with mysql 3.x # clients (those using the mysqlclient10 compatibility package). old_passwords=1 # Disabling symbolic-links is recommended to prevent assorted security risks ; # to do so, uncomment this line: # symbolic-links=0 [mysqld_safe] log-error=/var/log/mysqld.log pid-file=/var/run/mysqld/mysqld.pid bind-address=171.67.88.25 port=3306 And I have also restarted my firewall sudo /sbin/iptables -A INPUT -i eth0 -p tcp --destination-port 3306 -j ACCEPT /sbin/service iptables save The network administrator has already opened port 3306 for this box. When connecting from a remote computer (running Ubuntu 10.10, server is running RHEL 5.7), I issue mysql -u jeffrey -p --host=171.67.88.25 --port=3306 --socket=/var/lib/mysql/mysql.sock but receive a ERROR 2003 (HY000): Can't connect to MySQL server on '171.67.88.25' (113). I've noticed that the socket file /var/lib/mysql/mysql.sock is blank. Should this be the case? UPDATE The result of netstat -an | grep 3306 tcp 0 0 0.0.0.0:3306 0.0.0.0:* LISTEN Result of sudo netstat -tulpen Active Internet connections (only servers) Proto Recv-Q Send-Q Local Address Foreign Address State User Inode PID/Program name tcp 0 0 127.0.0.1:2208 0.0.0.0:* LISTEN 0 7602 3168/hpiod tcp 0 0 0.0.0.0:3306 0.0.0.0:* LISTEN 27 7827 3298/mysqld tcp 0 0 0.0.0.0:111 0.0.0.0:* LISTEN 0 5110 2802/portmap tcp 0 0 0.0.0.0:8787 0.0.0.0:* LISTEN 0 8431 3326/rserver tcp 0 0 0.0.0.0:915 0.0.0.0:* LISTEN 0 5312 2853/rpc.statd tcp 0 0 0.0.0.0:22 0.0.0.0:* LISTEN 0 7655 3188/sshd tcp 0 0 127.0.0.1:631 0.0.0.0:* LISTEN 0 7688 3199/cupsd tcp 0 0 127.0.0.1:25 0.0.0.0:* LISTEN 0 8025 3362/sendmail: acce tcp 0 0 127.0.0.1:2207 0.0.0.0:* LISTEN 0 7620 3173/python udp 0 0 0.0.0.0:909 0.0.0.0:* 0 5300 2853/rpc.statd udp 0 0 0.0.0.0:912 0.0.0.0:* 0 5309 2853/rpc.statd udp 0 0 0.0.0.0:68 0.0.0.0:* 0 4800 2598/dhclient udp 0 0 0.0.0.0:36177 0.0.0.0:* 70 8314 3476/avahi-daemon: udp 0 0 0.0.0.0:5353 0.0.0.0:* 70 8313 3476/avahi-daemon: udp 0 0 0.0.0.0:111 0.0.0.0:* 0 5109 2802/portmap udp 0 0 0.0.0.0:631 0.0.0.0:* 0 7691 3199/cupsd Result of sudo /sbin/iptables -L -v -n Chain INPUT (policy ACCEPT 0 packets, 0 bytes) pkts bytes target prot opt in out source destination 6373 2110K RH-Firewall-1-INPUT all -- * * 0.0.0.0/0 0.0.0.0/0 Chain FORWARD (policy ACCEPT 0 packets, 0 bytes) pkts bytes target prot opt in out source destination 0 0 RH-Firewall-1-INPUT all -- * * 0.0.0.0/0 0.0.0.0/0 Chain OUTPUT (policy ACCEPT 1241 packets, 932K bytes) pkts bytes target prot opt in out source destination Chain RH-Firewall-1-INPUT (2 references) pkts bytes target prot opt in out source destination 572 861K ACCEPT all -- lo * 0.0.0.0/0 0.0.0.0/0 1 28 ACCEPT icmp -- * * 0.0.0.0/0 0.0.0.0/0 icmp type 255 0 0 ACCEPT esp -- * * 0.0.0.0/0 0.0.0.0/0 0 0 ACCEPT ah -- * * 0.0.0.0/0 0.0.0.0/0 46 6457 ACCEPT udp -- * * 0.0.0.0/0 224.0.0.251 udp dpt:5353 0 0 ACCEPT udp -- * * 0.0.0.0/0 0.0.0.0/0 udp dpt:631 0 0 ACCEPT tcp -- * * 0.0.0.0/0 0.0.0.0/0 tcp dpt:631 782 157K ACCEPT all -- * * 0.0.0.0/0 0.0.0.0/0 state RELATED,ESTABLISHED 2 120 ACCEPT tcp -- * * 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:22 0 0 ACCEPT tcp -- * * 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:443 0 0 ACCEPT tcp -- * * 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:23 0 0 ACCEPT tcp -- * * 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:80 4970 1086K REJECT all -- * * 0.0.0.0/0 0.0.0.0/0 reject-with icmp-host-prohibited Result of nmap -P0 -p3306 171.67.88.25 Host is up (0.027s latency). PORT STATE SERVICE 3306/tcp filtered mysql Nmap done: 1 IP address (1 host up) scanned in 0.09 seconds Solution When everything else fails, go GUI! system-config-securitylevel and add port 3306. All done!

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  • Issues with Verizon's "Network Extender" device talking on my home network.

    - by Logan
    I recently switched my phone service to Verizon from ATT, and I get somewhat spotty service in my house. I called them and they sent me a "network extender" device for free. Its a femtocell that connects to my home network. The directions that come with it are very dumbed down, basically just say to connect it to your router and put it near a window (so it can get a GPS signal, it has to make sure its within the correct area before operating). The problem I'm having is the network light on it stays red. The troubleshooting information that came with it tells me this means there is a bad network connection. Its connected through an ASUS router running DD-WRT. No other devices on my network have a problem with it, including a Western Digital WDLIVE device, mine and my wife's cell phones (via wifi), a Wii, and an Xbox. If I connect the device directly to my cable modem, the light goes blue (which means good) and it starts working. So this tells me that its definately a configuration issue with my router. Verizon basically washed their hands of me when I connected it to my cable modem, and told me that its a router issue and to try a different router. Because normal people just have extra routers laying around their houses... When I connect it to the router, I can watch the DHCP Clients list on the status page, and the MAC of the network extender quickly fills up the clients list, grabbing every available DHCP address. Its like it grabs an address, can't connect to the internet, releases it, grabs another, then another, then another. So in the DHCP server settings I assigned a static IP to its MAC. This made it quit doing what it was doing before, but its still not working. I found the ports I needed to open on verizon's website, and opened them in the port forwarding config on my router. This still didn't help. So, I tried setting the network extender device's IP as the DMZ IP on the router. This still did no good. I called Verizon back and got the tech to write up a report which he passed on to a "senior network tech" who I got a call back from a few hours ago. This guy told me that while an ASUS router isn't listed as a supported device, he's not really sure why its not working. He suggested restoring the firmware to stock ASUS firmware and trying again. I have a very hard time believing its DD-WRT doing this, since every other device is working just fine with it. But its also not the Network Extender, since it works just fine when connected directly to the modem. At this point I'm out of ideas, and the next step is to restore the stock firmware on my router, and then going to walmart and getting a linksys WRT-54G to try. Is there anything else I could try before going that drastic? Cliffs- -Network extender won't work behind router, works when connected directly to cable modem. -Extender goes nuts when allowed to pick its own DHCP address, I had to assign it a static IP. -Won't work when correct ports are forwarded to it -Won't work with a DMZ address.

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  • Add a small RAID card? Will it help overall stability and performance of my nine hard drives?

    - by Ray
    Hi, Will I get any extra genuine added performance and RAID stability if I insert a basic RAID card into a PCI-E x1 slot? I am considering the Adaptec 1220SA - 2 port SATA , pci-express (1x) , raid 0/1. Ok it only supports two SATA drives. Purpose is to help support the eight internal hard drives (1TB each), a DVD drive and an external e-SATA connected 2TB hard drive - by dealing with two of the internal hard drives. My current configuration of eight internal 1TB Barracuda (7200.12) SATA hard drives, one external 2TB SATA Western Digital Green Drive (e-SATA) and one DVD drive can already be supported by the Intel P55 & JMicron controllers on the ASUS motherboard : the Intel P55 (controls six HDD; configured as three x RAID 1), and the JMicron (controls two HDD as one RAID 1, as well as the DVD drive and the external SATA drive via the motherboard's e-SATA port (controlled by the JMicron)). Bigger picture details : I have an ASUS motherboard designed for the LGA1156 type processor and it includes the Intel P55 Express Chipset and JMicron. I am using the Intel Core i7-870 processor, and have 8GB DDR3 (1333) memory (four x 2GB Corsair DIMMs). Enough overall power. The power supply is more than sufficicient for the system. Corsair AX850. The system will never need the full 850 watts (future : second graphics card). The RAID card would provide hardware RAID 1 for two of the eight intrnal drives. It would either reduce the load on : the Intel P55 firmware RAID support, or replace the JMicron controller's RAID 1 set. I am busy installing the above configuration using Windows 7 Ultimate 64-bit as the OS. The RAID card is a last minute addition to the plan. Is it worth spending the extra R700 - R900 on the Adaptec 1220SA, or equivalent RAID card? I cannot afford to spend yet another R2000 - R3000 on a RAID card that would support many SATA2 hard drives, with a better RAID, example the RAID 5. My Issue & assumption : I am trusting that the Intel P55 chipset can properly handle six drives, configured as three * RAID 1. I am assuming that the JMicron can handle, using its RED SATA ports, one RAID-1 (two HDDs). The DVD drive connects to the JMicron optical SATA port 1 (white port 1). White port 2 is not used. The e-SATA connection is from the JMicron straight to, and through the motherboard - to an on-board (rear panel) e-SATA port. Am I being a little hopeful in only using the on-board Intel P55 and the JMicron? Is it a waste of money to install a RAID card that handles two SATA2 drives? OR Is it wisdom to take the pressure a little off the Intel P55? Obviously I am interested in data security, hence RAID 1, not RAID Zero. RAID 5 would be nice. The CPU, Intel Core i7-870 will provide the clout. Context to nine drives : I am using virtualisation with Windows 7 Ultimate. Bootable VMs. The operating system gets a mirror. Loaded apps gets a mirror. The current design data is kept in another mirror and Another mirror is back-up one and / or VM territory. Then the external 2TB drive (via e-SATA) is the next layer of data security and then finally, I use off-site data security. Thanks.

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  • Troubleshooting sudoers via ldap

    - by dafydd
    The good news is that I got sudoers via ldap working on Red Hat Directory Server. The package is sudo-1.7.2p1. I have some LDAP/Kerberos users in an LDAP group called wheel, and I have this entry in LDAP: # %wheel, SUDOers, example.com dn: cn=%wheel,ou=SUDOers,dc=example,dc=com cn: %wheel description: Members of group wheel have access to all privileges. objectClass: sudoRole objectClass: top sudoCommand: ALL sudoHost: ALL sudoUser: %wheel So, members of group wheel have administrative privileges via sudo. This has been tested and works fine. Now, I have this other sudo privilege set up to allow members of a group called Administrators to perform two commands as the non-root owner of those commands. # %Administrators, SUDOers, example.com dn: cn=%Administrators,ou=SUDOers,dc=example,dc=com sudoRunAsGroup: appGroup sudoRunAsUser: appOwner cn: %Administrators description: Allow members of the group Administrators to run various commands . objectClass: sudoRole objectClass: top sudoCommand: appStop sudoCommand: appStart sudoCommand: /path/to/appStop sudoCommand: /path/to/appStart sudoUser: %Administrators Unfortunately, members of Administrators are still refused permission to run appStart or appStop: -bash-3.2$ sudo /path/to/appStop [sudo] password for Aaron: Sorry, user Aaron is not allowed to execute '/path/to/appStop' as root on host.example.com. -bash-3.2$ sudo -u appOwner /path/to/appStop [sudo] password for Aaron: Sorry, user Aaron is not allowed to execute '/path/to/appStop' as appOwner on host.example.com. /var/log/secure shows me these two sets of messages for the two attempts: Oct 31 15:02:36 host sudo: pam_unix(sudo:auth): authentication failure; logname=Aaron uid=0 euid=0 tty=/dev/pts/3 ruser= rhost= user=Aaron Oct 31 15:02:37 host sudo: pam_krb5[1508]: TGT verified using key for 'host/[email protected]' Oct 31 15:02:37 host sudo: pam_krb5[1508]: authentication succeeds for 'Aaron' ([email protected]) Oct 31 15:02:37 host sudo: Aaron : command not allowed ; TTY=pts/3 ; PWD=/auto/home/Aaron ; USER=root ; COMMAND=/path/to/appStop Oct 31 15:02:52 host sudo: pam_unix(sudo:auth): authentication failure; logname=Aaron uid=0 euid=0 tty=/dev/pts/3 ruser= rhost= user=Aaron Oct 31 15:02:52 host sudo: pam_krb5[1547]: TGT verified using key for 'host/[email protected]' Oct 31 15:02:52 host sudo: pam_krb5[1547]: authentication succeeds for 'Aaron' ([email protected]) Oct 31 15:02:52 host sudo: Aaron : command not allowed ; TTY=pts/3 ; PWD=/auto/home/Aaron ; USER=appOwner; COMMAND=/path/to/appStop The questions: Does sudo have some sort of verbose or debug mode where I can actually watch it capture the sudoers privilege list and determine whether or not Aaron should have the privilege to run this command? (This question is probably independent of where the sudoers database is kept.) Does sudo work with some background mechanism that might have a log level I could turn up? Right now, I can't fix a problem I can't identify. Is this an LDAP search failure? Is this a group member matching failure? Identifying why the command fails will help me identify the fix... Next step: Recreate the privilege in /etc/sudoers, and see if it works locally... Cheers!

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  • Nginx and multiple wordpress instances with fastcgi under same domain

    - by damnsweet
    My site is running on apache. two instances of wordpress exist under paths /tr/ and /eng/. I want to move the setup to nginx but could not manage to get it working. My setup consists of nging 0.7.66, php 5.3.2, and php-fpm. /tr/ and /eng/ are two separate wordpress instances located under /home/istci/webapps/wordpress_tr and /home/istci/webapps/wordpress respectively. Below is the server section from nginx.conf containing only configuration for tr, yet could not get it working either. server { listen 80; server_name www.example.com; charset utf-8; location ~ ^/$ { rewrite ^(.+)$ http://www.example.com/tr/ permanent; } location ~ /tr/.*php$ { fastcgi_pass unix:/home/istci/var/run/wptr.sock; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME /home/istci/webapps/wordpress_tr$fastcgi_script_name; fastcgi_param QUERY_STRING $query_string; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_param SCRIPT_NAME $fastcgi_script_name; fastcgi_param REQUEST_URI $request_uri; fastcgi_param DOCUMENT_URI $document_uri; fastcgi_param DOCUMENT_ROOT $document_root; fastcgi_param SERVER_PROTOCOL $server_protocol; fastcgi_param GATEWAY_INTERFACE CGI/1.1; fastcgi_param SERVER_SOFTWARE nginx/$nginx_version; fastcgi_param REMOTE_ADDR $remote_addr; fastcgi_param REMOTE_PORT $remote_port; fastcgi_param SERVER_ADDR $server_addr; fastcgi_param SERVER_PORT $server_port; fastcgi_param SERVER_NAME $server_name; # required if PHP was built with --enable-force-cgi-redirect fastcgi_param REDIRECT_STATUS 200; } location /tr/ { root /home/istci/webapps/wordpress_tr/; index index.php index.html index.htm; if (!-e $request_filename) { rewrite ^(.+)$ /tr/index.php?q=$1 last; break; } if (-f $request_filename) { expires 30d; break; } } } php-fpm listens on unix:/home/istci/var/run/wptr.sock. running it in debug-mode shows no active handlers, which means no connection is made to unix socket from nginx. nginx access logs: 127.0.0.1 - - [09/Jun/2010:03:45:11 -0500] "GET /tr/ HTTP/1.0" 404 20 "-" "Mozilla/5.0 (X11; U; Linux i686; en-US; rv:1.9.2.4) Gecko/20100527 Firefox/3.6.4" nginx debug logs : 2010/06/09 03:38:53 [notice] 6922#0: built by gcc 4.1.2 20080704 (Red Hat 4.1.2-48) 2010/06/09 03:38:53 [notice] 6922#0: OS: Linux 2.6.18-164.9.1.el5PAE 2010/06/09 03:38:53 [notice] 6922#0: getrlimit(RLIMIT_NOFILE): 4096:4096 2010/06/09 03:38:53 [notice] 6923#0: start worker processes 2010/06/09 03:38:53 [notice] 6923#0: start worker process 6924 2010/06/09 03:38:53 [notice] 6923#0: start worker process 6925 2010/06/09 03:39:01 [notice] 6925#0: *1 "^(.+)$" matches "/tr/", client: 127.0.0.1, server: www.example.com, request: "GET /tr/ HTTP/1.0", host: "www.example.com" 2010/06/09 03:39:01 [notice] 6925#0: *1 rewritten data: "/tr/index.php", args: "q=/tr/", client: 127.0.0.1, server: www.example.com, request: "GET /tr/ HTTP/1.0", host: "www.example.com" Any clues about what is wrong with my configuration? Thanks.

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  • redhat Apache fast-cgi selinux permissions

    - by Alejo JM
    My apache installation is running php as fastcgi, and the virtual hosts are pointing to /home/*/public_html. and the fastcgi are home/*/cgi-bin/php.fcgi the public_html setup with selinux was: /usr/sbin/setsebool -P httpd_enable_homedirs 1 chcon -R -t httpd_sys_content_t /home/someuser/public_html The owner and group are the user, for example the user "someuser": ls -all /home/someuser/cgi-bin/ drwxr-xr-x 2 someuser someuser 4096 Sep 7 13:14 . drwx--x--x 6 someuser someuser 4096 Sep 6 18:17 .. -rwxr-xr-x 1 someuser someuser 308 Sep 7 13:14 php.fcgi ls -all /home/someuser/public_html/ | greep info.php -rw-r--r-- 1 someuser someuser 24 Sep 3 16:24 info.php When is visits the site I get "Forbidden ..." and the log said: [Fri Sep 07 12:02:51 2012] [error] [client x.x.x.x] (13)Permission denied: access to /cgi-bin/php.fcgi/info.php denied My selinux conf is: SELINUX=enforcing SELINUXTYPE=targeted SETLOCALDEFS=0 So I kill Selinux (SELINUX=disabled), reboot the system and everything works !!!!! The problem is Selinux, I don't want disable Selinux. I trying this with no success: setsebool -P httpd_enable_cgi 1 chcon -t httpd_sys_script_exec_t /home/someuser/cgi-bin/php.fcgi chcon -R -t httpd_sys_content_t /home/someuser/cgi-bin Or maybe is better change Selinux SELINUX=enforcing to SELINUX=permissive And disable selinux for httpd ? (I think I better find the correct configuration) Thanks for any suggestion on this matter My environment: Red Hat Enterprise Linux Server release 5.8 (Tikanga) Server version: Apache/2.2.3 PHP 5.1.6 (cli) (built: Jun 22 2012 06:20:25) Copyright (c) 1997-2006 The PHP Group Zend Engine v2.1.0, Copyright (c) 1998-2006 Zend Technologies Some logs: ps -ZC httpd LABEL PID TTY TIME CMD system_u:system_r:httpd_t 2822 ? 00:00:00 httpd system_u:system_r:httpd_t 2823 ? 00:00:00 httpd system_u:system_r:httpd_t 2824 ? 00:00:00 httpd system_u:system_r:httpd_t 2825 ? 00:00:00 httpd system_u:system_r:httpd_t 2826 ? 00:00:00 httpd system_u:system_r:httpd_t 2836 ? 00:00:00 httpd system_u:system_r:httpd_t 2837 ? 00:00:00 httpd system_u:system_r:httpd_t 2838 ? 00:00:00 httpd system_u:system_r:httpd_t 2839 ? 00:00:00 httpd system_u:system_r:httpd_t 2840 ? 00:00:00 httpd getsebool -a | grep httpd allow_httpd_anon_write --> off allow_httpd_bugzilla_script_anon_write --> off allow_httpd_cvs_script_anon_write --> off allow_httpd_mod_auth_pam --> off allow_httpd_nagios_script_anon_write --> off allow_httpd_prewikka_script_anon_write --> off allow_httpd_squid_script_anon_write --> off allow_httpd_sys_script_anon_write --> off httpd_builtin_scripting --> on httpd_can_network_connect --> off httpd_can_network_connect_db --> off httpd_can_network_relay --> off httpd_can_sendmail --> on httpd_disable_trans --> off httpd_enable_cgi --> on httpd_enable_ftp_server --> off httpd_enable_homedirs --> on httpd_execmem --> off httpd_read_user_content --> off httpd_rotatelogs_disable_trans --> off httpd_setrlimit --> off httpd_ssi_exec --> off httpd_suexec_disable_trans --> off httpd_tty_comm --> on httpd_unified --> on httpd_use_cifs --> off httpd_use_nfs --> off

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  • dns server bind is not work

    - by milad
    I just installed bind on RHEL 6 and point a domain to that server. but actually when i ping domain it returns error 1214: Here is my named.conf: // // named.conf // // Provided by Red Hat bind package to configure the ISC BIND named(8) DNS // server as a caching only nameserver (as a localhost DNS resolver only). // // See /usr/share/doc/bind*/sample/ for example named configuration files. // options { listen-on port 53 { any; }; listen-on-v6 port 53 { ::1; }; directory "/var/named"; dump-file "/var/named/data/cache_dump.db"; statistics-file "/var/named/data/named_stats.txt"; memstatistics-file "/var/named/data/named_mem_stats.txt"; allow-query { any; }; recursion yes; dnssec-enable yes; dnssec-validation yes; dnssec-lookaside auto; /* Path to ISC DLV key */ bindkeys-file "/etc/named.iscdlv.key"; managed-keys-directory "/var/named/dynamic"; }; logging { channel default_debug { file "data/named.run"; severity dynamic; }; }; zone "." IN { type hint; file "named.ca"; }; include "/etc/named.rfc1912.zones"; include "/etc/named.root.key"; zone "mydomain.com"{ type master; file "/var/named/data/named.mydomain.com"; allow-update { none; }; };` AND The content of "/var/named/data/named.mydomain.com": $TTL 38400 mydomain.com. IN SOA ns1.mydomain.com. milad.yahoo.com. ( 2012101201 ; serial number YYMMDDNN 28800 ; Refresh 7200 ; Retry 864000 ; Expire 38400 ; Min TTL ) mydomain.com. IN A 1.2.3.4 www IN A 1.2.3.4 ns1.mydomain.com. IN A 1.2.3.4 ns2.mydomain.com. IN A 1.2.3.4 mydomain.com. IN NS ns1.mydomain.com. mydomain.com. IN NS ns2.mydomain.com. AND i'm sure the named service is running: [root@server ~]# service named status version: 9.8.2rc1-RedHat-9.8.2-0.10.rc1.el6_3.3 CPUs found: 8 worker threads: 8 number of zones: 20 debug level: 0 xfers running: 0 xfers deferred: 0 soa queries in progress: 0 query logging is OFF recursive clients: 0/0/1000 tcp clients: 0/100 server is up and running named (pid 26299) is running... Thanks for your answers. i know that the ping is not the job of bind, i use it just to check whether domain is pointed to host or not.(ping is open in my server as i got reply in pinging ip) i use network-tools.com to ping domain. here the output of dig utility: dig mydomain.com ; <<>> DiG 9.8.2rc1-RedHat-9.8.2-0.10.rc1.el6_3.3 <<>> mydomain.com ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: SERVFAIL, id: 6806 ;; flags: qr rd ra; QUERY: 1, ANSWER: 0, AUTHORITY: 0, ADDITIONAL: 0 ;; QUESTION SECTION: ;mydomain.com. IN A ;; Query time: 321 msec ;; SERVER: 5.6.7.8#53(5.6.7.8)##note that 5.6.7.8 is my idc dns ip ;; WHEN: Sun Oct 14 23:53:47 2012

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  • Call to daemon in a /etc/init.d script is blocking, not running in background

    - by tony
    I have a Perl script that I want to daemonize. Basically this perl script will read a directory every 30 seconds, read the files that it finds and then process the data. To keep it simple here consider the following Perl script (called synpipe_server, there is a symbolic link of this script in /usr/sbin/) : #!/usr/bin/perl use strict; use warnings; my $continue = 1; $SIG{'TERM'} = sub { $continue = 0; print "Caught TERM signal\n"; }; $SIG{'INT'} = sub { $continue = 0; print "Caught INT signal\n"; }; my $i = 0; while ($continue) { #do stuff print "Hello, I am running " . ++$i . "\n"; sleep 3; } So this script basically prints something every 3 seconds. Then, as I want to daemonize this script, I've also put this bash script (also called synpipe_server) in /etc/init.d/ : #!/bin/bash # synpipe_server : This starts and stops synpipe_server # # chkconfig: 12345 12 88 # description: Monitors all production pipelines # processname: synpipe_server # pidfile: /var/run/synpipe_server.pid # Source function library. . /etc/rc.d/init.d/functions pname="synpipe_server" exe="/usr/sbin/synpipe_server" pidfile="/var/run/${pname}.pid" lockfile="/var/lock/subsys/${pname}" [ -x $exe ] || exit 0 RETVAL=0 start() { echo -n "Starting $pname : " daemon ${exe} RETVAL=$? PID=$! echo [ $RETVAL -eq 0 ] && touch ${lockfile} echo $PID > ${pidfile} } stop() { echo -n "Shutting down $pname : " killproc ${exe} RETVAL=$? echo if [ $RETVAL -eq 0 ]; then rm -f ${lockfile} rm -f ${pidfile} fi } restart() { echo -n "Restarting $pname : " stop sleep 2 start } case "$1" in start) start ;; stop) stop ;; status) status ${pname} ;; restart) restart ;; *) echo "Usage: $0 {start|stop|status|restart}" ;; esac exit 0 So, (if I have well understood the doc for daemon) the Perl script should run in the background and the output should be redirected to /dev/null if I execute : service synpipe_server start But here is what I get instead : [root@master init.d]# service synpipe_server start Starting synpipe_server : Hello, I am running 1 Hello, I am running 2 Hello, I am running 3 Hello, I am running 4 Caught INT signal [ OK ] [root@master init.d]# So it starts the Perl script but runs it without detaching it from the current terminal session, and I can see the output printed in my console ... which is not really what I was expecting. Moreover, the PID file is empty (or with a line feed only, no pid returned by daemon). Does anyone have any idea of what I am doing wrong ? EDIT : maybe I should say that I am on a Red Hat machine. Scientific Linux SL release 5.4 (Boron) Would it do the job if instead of using the daemon function, I use something like : nohup ${exe} >/dev/null 2>&1 & in the init script ?

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