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  • Is it possible to write syntax like - ()() ?

    - by dotnetdev
    I read in an ebook somewhere (which I'm desperate to find again), that, by using delegates, it is possible to write code which has syntax as follows: ()(); // where delegate proceeds this. Can anyone provide any details how this would be possible/in what situation this would occur? Thanks

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  • Saving an Object for use later

    - by Eclipsed4utoo
    As part of my widget, I use an instance of the Camera object. This is what I want to do. The user will click on my widget, I get an instance of the Camera(if it's not already stored), use it, then store it. If they click the widget again, I want to use that same instance that I used previously. Is this possible?

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  • How remove node ID [nid:n] in NodeReference fields

    - by Snazzy
    Hi. This is the same question of this link: http://stackoverflow.com/questions/1515722/removing-nidn-in-nodereference-autocomplete According with the first answer (Grayside) I've created my own module and activated. Then I create a new content, I look sth up in the nodereference field and finally select it - it works (Doesn't appear the [nid:n]). But, when I view/preview or save or edit the content, the [nid:n] appears again. Anybody can help me?

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  • How to Stop the Window Animation in Win XP SP3 Permanently??

    - by epale
    Hi everyone, May I know how I can get rid of the Window Animation (seen when you minimise or maximise a window) in Win XP SP3 Permanently?? I have tried using windows powertoys tweakUI as well as going to control panel---adjust visual effects--- then unchecking the "Animate windows when maximising and minimising" option. Problem is that the window animation will disappear at first but returns again some time later. Thank you very much

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  • Building applications with WPF, MVVM and Prism(aka CAG)

    - by skjagini
    In this article I am going to walk through an application using WPF and Prism (aka composite application guidance, CAG) which simulates engaging a taxi (cab).  The rules are simple, the app would have3 screens A login screen to authenticate the user An information screen. A screen to engage the cab and roam around and calculating the total fare Metered Rate of Fare The meter is required to be engaged when a cab is occupied by anyone $3.00 upon entry $0.35 for each additional unit The unit fare is: one-fifth of a mile, when the cab is traveling at 6 miles an hour or more; or 60 seconds when not in motion or traveling at less than 12 miles per hour. Night surcharge of $.50 after 8:00 PM & before 6:00 AM Peak hour Weekday Surcharge of $1.00 Monday - Friday after 4:00 PM & before 8:00 PM New York State Tax Surcharge of $.50 per ride. Example: Friday (2010-10-08) 5:30pm Start at Lexington Ave & E 57th St End at Irving Pl & E 15th St Start = $3.00 Travels 2 miles at less than 6 mph for 15 minutes = $3.50 Travels at more than 12 mph for 5 minutes = $1.75 Peak hour Weekday Surcharge = $1.00 (ride started at 5:30 pm) New York State Tax Surcharge = $0.50 Before we dive into the app, I would like to give brief description about the framework.  If you want to jump on to the source code, scroll all the way to the end of the post. MVVM MVVM pattern is in no way related to the usage of PRISM in your application and should be considered if you are using WPF irrespective of PRISM or not. Lets say you are not familiar with MVVM, your typical UI would involve adding some UI controls like text boxes, a button, double clicking on the button,  generating event handler, calling a method from business layer and updating the user interface, it works most of the time for developing small scale applications. The problem with this approach is that there is some amount of code specific to business logic wrapped in UI specific code which is hard to unit test it, mock it and MVVM helps to solve the exact problem. MVVM stands for Model(M) – View(V) – ViewModel(VM),  based on the interactions with in the three parties it should be called VVMM,  MVVM sounds more like MVC (Model-View-Controller) so the name. Why it should be called VVMM: View – View Model - Model WPF allows to create user interfaces using XAML and MVVM takes it to the next level by allowing complete separation of user interface and business logic. In WPF each view will have a property, DataContext when set to an instance of a class (which happens to be your view model) provides the data the view is interested in, i.e., view interacts with view model and at the same time view model interacts with view through DataContext. Sujith, if view and view model are interacting directly with each other how does MVVM is helping me separation of concerns? Well, the catch is DataContext is of type Object, since it is of type object view doesn’t know exact type of view model allowing views and views models to be loosely coupled. View models aggregate data from models (data access layer, services, etc) and make it available for views through properties, methods etc, i.e., View Models interact with Models. PRISM Prism is provided by Microsoft Patterns and Practices team and it can be downloaded from codeplex for source code,  samples and documentation on msdn.  The name composite implies, to compose user interface from different modules (views) without direct dependencies on each other, again allowing  loosely coupled development. Well Sujith, I can already do that with user controls, why shall I learn another framework?  That’s correct, you can decouple using user controls, but you still have to manage some amount of coupling, like how to do you communicate between the controls, how do you subscribe/unsubscribe, loading/unloading views dynamically. Prism is not a replacement for user controls, provides the following features which greatly help in designing the composite applications. Dependency Injection (DI)/ Inversion of Control (IoC) Modules Regions Event Aggregator  Commands Simply put, MVVM helps building a single view and Prism helps building an application using the views There are other open source alternatives to Prism, like MVVMLight, Cinch, take a look at them as well. Lets dig into the source code.  1. Solution The solution is made of the following projects Framework: Holds the common functionality in building applications using WPF and Prism TaxiClient: Start up project, boot strapping and app styling TaxiCommon: Helps with the business logic TaxiModules: Holds the meat of the application with views and view models TaxiTests: To test the application 2. DI / IoC Dependency Injection (DI) as the name implies refers to injecting dependencies and Inversion of Control (IoC) means the calling code has no direct control on the dependencies, opposite of normal way of programming where dependencies are passed by caller, i.e inversion; aside from some differences in terminology the concept is same in both the cases. The idea behind DI/IoC pattern is to reduce the amount of direct coupling between different components of the application, the higher the dependency the more tightly coupled the application resulting in code which is hard to modify, unit test and mock.  Initializing Dependency Injection through BootStrapper TaxiClient is the starting project of the solution and App (App.xaml)  is the starting class that gets called when you run the application. From the App’s OnStartup method we will invoke BootStrapper.   namespace TaxiClient { /// <summary> /// Interaction logic for App.xaml /// </summary> public partial class App : Application { protected override void OnStartup(StartupEventArgs e) { base.OnStartup(e);   (new BootStrapper()).Run(); } } } BootStrapper is your contact point for initializing the application including dependency injection, creating Shell and other frameworks. We are going to use Unity for DI and there are lot of open source DI frameworks like Spring.Net, StructureMap etc with different feature set  and you can choose a framework based on your preferences. Note that Prism comes with in built support for Unity, for example we are deriving from UnityBootStrapper in our case and for any other DI framework you have to extend the Prism appropriately   namespace TaxiClient { public class BootStrapper: UnityBootstrapper { protected override IModuleCatalog CreateModuleCatalog() { return new ConfigurationModuleCatalog(); } protected override DependencyObject CreateShell() { Framework.FrameworkBootStrapper.Run(Container, Application.Current.Dispatcher);   Shell shell = new Shell(); shell.ResizeMode = ResizeMode.NoResize; shell.Show();   return shell; } } } Lets take a look into  FrameworkBootStrapper to check out how to register with unity container. namespace Framework { public class FrameworkBootStrapper { public static void Run(IUnityContainer container, Dispatcher dispatcher) { UIDispatcher uiDispatcher = new UIDispatcher(dispatcher); container.RegisterInstance<IDispatcherService>(uiDispatcher);   container.RegisterType<IInjectSingleViewService, InjectSingleViewService>( new ContainerControlledLifetimeManager());   . . . } } } In the above code we are registering two components with unity container. You shall observe that we are following two different approaches, RegisterInstance and RegisterType.  With RegisterInstance we are registering an existing instance and the same instance will be returned for every request made for IDispatcherService   and with RegisterType we are requesting unity container to create an instance for us when required, i.e., when I request for an instance for IInjectSingleViewService, unity will create/return an instance of InjectSingleViewService class and with RegisterType we can configure the life time of the instance being created. With ContaienrControllerLifetimeManager, the unity container caches the instance and reuses for any subsequent requests, without recreating a new instance. Lets take a look into FareViewModel.cs and it’s constructor. The constructor takes one parameter IEventAggregator and if you try to find all references in your solution for IEventAggregator, you will not find a single location where an instance of EventAggregator is passed directly to the constructor. The compiler still finds an instance and works fine because Prism is already configured when used with Unity container to return an instance of EventAggregator when requested for IEventAggregator and in this particular case it is called constructor injection. public class FareViewModel:ObservableBase, IDataErrorInfo { ... private IEventAggregator _eventAggregator;   public FareViewModel(IEventAggregator eventAggregator) { _eventAggregator = eventAggregator; InitializePropertyNames(); InitializeModel(); PropertyChanged += OnPropertyChanged; } ... 3. Shell Shells are very similar in operation to Master Pages in asp.net or MDI in Windows Forms. And shells contain regions which display the views, you can have as many regions as you wish in a given view. You can also nest regions. i.e, one region can load a view which in itself may contain other regions. We have to create a shell at the start of the application and are doing it by overriding CreateShell method from BootStrapper From the following Shell.xaml you shall notice that we have two content controls with Region names as ‘MenuRegion’ and ‘MainRegion’.  The idea here is that you can inject any user controls into the regions dynamically, i.e., a Menu User Control for MenuRegion and based on the user action you can load appropriate view into MainRegion.    <Window x:Class="TaxiClient.Shell" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:Regions="clr-namespace:Microsoft.Practices.Prism.Regions;assembly=Microsoft.Practices.Prism" Title="Taxi" Height="370" Width="800"> <Grid Margin="2"> <ContentControl Regions:RegionManager.RegionName="MenuRegion" HorizontalAlignment="Stretch" VerticalAlignment="Stretch" HorizontalContentAlignment="Stretch" VerticalContentAlignment="Stretch" />   <ContentControl Grid.Row="1" Regions:RegionManager.RegionName="MainRegion" HorizontalAlignment="Stretch" VerticalAlignment="Stretch" HorizontalContentAlignment="Stretch" VerticalContentAlignment="Stretch" /> <!--<Border Grid.ColumnSpan="2" BorderThickness="2" CornerRadius="3" BorderBrush="LightBlue" />-->   </Grid> </Window> 4. Modules Prism provides the ability to build composite applications and modules play an important role in it. For example if you are building a Mortgage Loan Processor application with 3 components, i.e. customer’s credit history,  existing mortgages, new home/loan information; and consider that the customer’s credit history component involves gathering data about his/her address, background information, job details etc. The idea here using Prism modules is to separate the implementation of these 3 components into their own visual studio projects allowing to build components with no dependency on each other and independently. If we need to add another component to the application, the component can be developed by in house team or some other team in the organization by starting with a new Visual Studio project and adding to the solution at the run time with very little knowledge about the application. Prism modules are defined by implementing the IModule interface and each visual studio project to be considered as a module should implement the IModule interface.  From the BootStrapper.cs you shall observe that we are overriding the method by returning a ConfiguratingModuleCatalog which returns the modules that are registered for the application using the app.config file  and you can also add module using code. Lets take a look into configuration file.   <?xml version="1.0"?> <configuration> <configSections> <section name="modules" type="Microsoft.Practices.Prism.Modularity.ModulesConfigurationSection, Microsoft.Practices.Prism"/> </configSections> <modules> <module assemblyFile="TaxiModules.dll" moduleType="TaxiModules.ModuleInitializer, TaxiModules" moduleName="TaxiModules"/> </modules> </configuration> Here we are adding TaxiModules project to our solution and TaxiModules.ModuleInitializer implements IModule interface   5. Module Mapper With Prism modules you can dynamically add or remove modules from the regions, apart from that Prism also provides API to control adding/removing the views from a region within the same module. Taxi Information Screen: Engage the Taxi Screen: The sample application has two screens, ‘Taxi Information’ and ‘Engage the Taxi’ and they both reside in same module, TaxiModules. ‘Engage the Taxi’ is again made of two user controls, FareView on the left and TotalView on the right. We have created a Shell with two regions, MenuRegion and MainRegion with menu loaded into MenuRegion. We can create a wrapper user control called EngageTheTaxi made of FareView and TotalView and load either TaxiInfo or EngageTheTaxi into MainRegion based on the user action. Though it will work it tightly binds the user controls and for every combination of user controls, we need to create a dummy wrapper control to contain them. Instead we can apply the principles we learned so far from Shell/regions and introduce another template (LeftAndRightRegionView.xaml) made of two regions Region1 (left) and Region2 (right) and load  FareView and TotalView dynamically.  To help with loading of the views dynamically I have introduce an helper an interface, IInjectSingleViewService,  idea suggested by Mike Taulty, a must read blog for .Net developers. using System; using System.Collections.Generic; using System.ComponentModel;   namespace Framework.PresentationUtility.Navigation {   public interface IInjectSingleViewService : INotifyPropertyChanged { IEnumerable<CommandViewDefinition> Commands { get; } IEnumerable<ModuleViewDefinition> Modules { get; }   void RegisterViewForRegion(string commandName, string viewName, string regionName, Type viewType); void ClearViewFromRegion(string viewName, string regionName); void RegisterModule(string moduleName, IList<ModuleMapper> moduleMappers); } } The Interface declares three methods to work with views: RegisterViewForRegion: Registers a view with a particular region. You can register multiple views and their regions under one command.  When this particular command is invoked all the views registered under it will be loaded into their regions. ClearViewFromRegion: To unload a specific view from a region. RegisterModule: The idea is when a command is invoked you can load the UI with set of controls in their default position and based on the user interaction, you can load different contols in to different regions on the fly.  And it is supported ModuleViewDefinition and ModuleMappers as shown below. namespace Framework.PresentationUtility.Navigation { public class ModuleViewDefinition { public string ModuleName { get; set; } public IList<ModuleMapper> ModuleMappers; public ICommand Command { get; set; } }   public class ModuleMapper { public string ViewName { get; set; } public string RegionName { get; set; } public Type ViewType { get; set; } } } 6. Event Aggregator Prism event aggregator enables messaging between components as in Observable pattern, Notifier notifies the Observer which receives notification it is interested in. When it comes to Observable pattern, Observer has to unsubscribes for notifications when it no longer interested in notifications, which allows the Notifier to remove the Observer’s reference from it’s local cache. Though .Net has managed garbage collection it cannot remove inactive the instances referenced by an active instance resulting in memory leak, keeping the Observers in memory as long as Notifier stays in memory.  Developers have to be very careful to unsubscribe when necessary and it often gets overlooked, to overcome these problems Prism Event Aggregator uses weak references to cache the reference (Observer in this case)  and releases the reference (memory) once the instance goes out of scope. Using event aggregator is very simple, declare a generic type of CompositePresenationEvent by inheriting from it. using Microsoft.Practices.Prism.Events; using TaxiCommon.BAO;   namespace TaxiCommon.CompositeEvents { public class TaxiOnMoveEvent:CompositePresentationEvent<TaxiOnMove> { } }   TaxiOnMove.cs includes the properties which we want to exchange between the parties, FareView and TotalView. using System;   namespace TaxiCommon.BAO { public class TaxiOnMove { public TimeSpan MinutesAtTweleveMPH { get; set; } public double MilesAtSixMPH { get; set; } } }   Lets take a look into FareViewodel (Notifier) and how it raises the event.  Here we are raising the event by getting the event through GetEvent<..>() and publishing it with the payload private void OnAddMinutes(object obj) { TaxiOnMove payload = new TaxiOnMove(); if(MilesAtSixMPH != null) payload.MilesAtSixMPH = MilesAtSixMPH.Value; if(MinutesAtTweleveMPH != null) payload.MinutesAtTweleveMPH = new TimeSpan(0,0,MinutesAtTweleveMPH.Value,0);   _eventAggregator.GetEvent<TaxiOnMoveEvent>().Publish(payload); ResetMinutesAndMiles(); } And TotalViewModel(Observer) subscribes to notifications by getting the event through GetEvent<..>() namespace TaxiModules.ViewModels { public class TotalViewModel:ObservableBase { .... private IEventAggregator _eventAggregator;   public TotalViewModel(IEventAggregator eventAggregator) { _eventAggregator = eventAggregator; ... }   private void SubscribeToEvents() { _eventAggregator.GetEvent<TaxiStartedEvent>() .Subscribe(OnTaxiStarted, ThreadOption.UIThread,false,(filter) => true); _eventAggregator.GetEvent<TaxiOnMoveEvent>() .Subscribe(OnTaxiMove, ThreadOption.UIThread, false, (filter) => true); _eventAggregator.GetEvent<TaxiResetEvent>() .Subscribe(OnTaxiReset, ThreadOption.UIThread, false, (filter) => true); }   ... private void OnTaxiMove(TaxiOnMove taxiOnMove) { OnMoveFare fare = new OnMoveFare(taxiOnMove); Fares.Add(fare); SetTotalFare(new []{fare}); }   .... 7. MVVM through example In this section we are going to look into MVVM implementation through example.  I have all the modules declared in a single project, TaxiModules, again it is not necessary to have them into one project. Once the user logs into the application, will be greeted with the ‘Engage the Taxi’ screen which is made of two user controls, FareView.xaml and TotalView.Xaml. As you can see from the solution explorer, each of them have their own code behind files and  ViewModel classes, FareViewMode.cs, TotalViewModel.cs Lets take a look in to the FareView and how it interacts with FareViewModel using MVVM implementation. FareView.xaml acts as a view and FareViewMode.cs is it’s view model. The FareView code behind class   namespace TaxiModules.Views { /// <summary> /// Interaction logic for FareView.xaml /// </summary> public partial class FareView : UserControl { public FareView(FareViewModel viewModel) { InitializeComponent(); this.Loaded += (s, e) => { this.DataContext = viewModel; }; } } } The FareView is bound to FareViewModel through the data context  and you shall observe that DataContext is of type Object, i.e. the FareView doesn’t really know the type of ViewModel (FareViewModel). This helps separation of View and ViewModel as View and ViewModel are independent of each other, you can bind FareView to FareViewModel2 as well and the application compiles just fine. Lets take a look into FareView xaml file  <UserControl x:Class="TaxiModules.Views.FareView" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:Toolkit="clr-namespace:Microsoft.Windows.Controls;assembly=WPFToolkit" xmlns:Commands="clr-namespace:Microsoft.Practices.Prism.Commands;assembly=Microsoft.Practices.Prism"> <Grid Margin="10" > ....   <Border Style="{DynamicResource innerBorder}" Grid.Row="0" Grid.Column="0" Grid.RowSpan="11" Grid.ColumnSpan="2" Panel.ZIndex="1"/>   <Label Grid.Row="0" Content="Engage the Taxi" Style="{DynamicResource innerHeader}"/> <Label Grid.Row="1" Content="Select the State"/> <ComboBox Grid.Row="1" Grid.Column="1" ItemsSource="{Binding States}" Height="auto"> <ComboBox.ItemTemplate> <DataTemplate> <TextBlock Text="{Binding Name}"/> </DataTemplate> </ComboBox.ItemTemplate> <ComboBox.SelectedItem> <Binding Path="SelectedState" Mode="TwoWay"/> </ComboBox.SelectedItem> </ComboBox> <Label Grid.Row="2" Content="Select the Date of Entry"/> <Toolkit:DatePicker Grid.Row="2" Grid.Column="1" SelectedDate="{Binding DateOfEntry, ValidatesOnDataErrors=true}" /> <Label Grid.Row="3" Content="Enter time 24hr format"/> <TextBox Grid.Row="3" Grid.Column="1" Text="{Binding TimeOfEntry, TargetNullValue=''}"/> <Button Grid.Row="4" Grid.Column="1" Content="Start the Meter" Commands:Click.Command="{Binding StartMeterCommand}" />   <Label Grid.Row="5" Content="Run the Taxi" Style="{DynamicResource innerHeader}"/> <Label Grid.Row="6" Content="Number of Miles &lt;@6mph"/> <TextBox Grid.Row="6" Grid.Column="1" Text="{Binding MilesAtSixMPH, TargetNullValue='', ValidatesOnDataErrors=true}"/> <Label Grid.Row="7" Content="Number of Minutes @12mph"/> <TextBox Grid.Row="7" Grid.Column="1" Text="{Binding MinutesAtTweleveMPH, TargetNullValue=''}"/> <Button Grid.Row="8" Grid.Column="1" Content="Add Minutes and Miles " Commands:Click.Command="{Binding AddMinutesCommand}"/> <Label Grid.Row="9" Content="Other Operations" Style="{DynamicResource innerHeader}"/> <Button Grid.Row="10" Grid.Column="1" Content="Reset the Meter" Commands:Click.Command="{Binding ResetCommand}"/>   </Grid> </UserControl> The highlighted code from the above code shows data binding, for example ComboBox which displays list of states has it’s ItemsSource bound to States property, with DataTemplate bound to Name and SelectedItem  to SelectedState. You might be wondering what are all these properties and how it is able to bind to them.  The answer lies in data context, i.e., when you bound a control, WPF looks for data context on the root object (Grid in this case) and if it can’t find data context it will look into root’s root, i.e. FareView UserControl and it is bound to FareViewModel.  Each of those properties have be declared on the ViewModel for the View to bind correctly. To put simply, View is bound to ViewModel through data context of type object and every control that is bound on the View actually binds to the public property on the ViewModel. Lets look into the ViewModel code (the following code is not an exact copy of FareViewMode.cs, pasted relevant code for this section)   namespace TaxiModules.ViewModels { public class FareViewModel:ObservableBase, IDataErrorInfo { public List<USState> States { get { return USStates.StateList; } }   public USState SelectedState { get { return _selectedState; } set { _selectedState = value; RaisePropertyChanged(_selectedStatePropertyName); } }   public DateTime? DateOfEntry { get { return _dateOfEntry; } set { _dateOfEntry = value; RaisePropertyChanged(_dateOfEntryPropertyName); } }   public TimeSpan? TimeOfEntry { get { return _timeOfEntry; } set { _timeOfEntry = value; RaisePropertyChanged(_timeOfEntryPropertyName); } }   public double? MilesAtSixMPH { get { return _milesAtSixMPH; } set { _milesAtSixMPH = value; RaisePropertyChanged(_distanceAtSixMPHPropertyName); } }   public int? MinutesAtTweleveMPH { get { return _minutesAtTweleveMPH; } set { _minutesAtTweleveMPH = value; RaisePropertyChanged(_minutesAtTweleveMPHPropertyName); } }   public ICommand StartMeterCommand { get { if(_startMeterCommand == null) { _startMeterCommand = new DelegateCommand<object>(OnStartMeter, CanStartMeter); } return _startMeterCommand; } }   public ICommand AddMinutesCommand { get { if(_addMinutesCommand == null) { _addMinutesCommand = new DelegateCommand<object>(OnAddMinutes, CanAddMinutes); } return _addMinutesCommand; } }   public ICommand ResetCommand { get { if(_resetCommand == null) { _resetCommand = new DelegateCommand<object>(OnResetCommand); } return _resetCommand; } }   } private void OnStartMeter(object obj) { _eventAggregator.GetEvent<TaxiStartedEvent>().Publish( new TaxiStarted() { EngagedOn = DateOfEntry.Value.Date + TimeOfEntry.Value, EngagedState = SelectedState.Value });   _isMeterStarted = true; OnPropertyChanged(this,null); } And views communicate user actions like button clicks, tree view item selections, etc using commands. When user clicks on ‘Start the Meter’ button it invokes the method StartMeterCommand, which calls the method OnStartMeter which publishes the event to TotalViewModel using event aggregator  and TaxiStartedEvent. namespace TaxiModules.ViewModels { public class TotalViewModel:ObservableBase { ... private IEventAggregator _eventAggregator;   public TotalViewModel(IEventAggregator eventAggregator) { _eventAggregator = eventAggregator;   InitializePropertyNames(); InitializeModel(); SubscribeToEvents(); }   public decimal? TotalFare { get { return _totalFare; } set { _totalFare = value; RaisePropertyChanged(_totalFarePropertyName); } } .... private void SubscribeToEvents() { _eventAggregator.GetEvent<TaxiStartedEvent>().Subscribe(OnTaxiStarted, ThreadOption.UIThread,false,(filter) => true); _eventAggregator.GetEvent<TaxiOnMoveEvent>().Subscribe(OnTaxiMove, ThreadOption.UIThread, false, (filter) => true); _eventAggregator.GetEvent<TaxiResetEvent>().Subscribe(OnTaxiReset, ThreadOption.UIThread, false, (filter) => true); }   private void OnTaxiStarted(TaxiStarted taxiStarted) { Fares.Add(new EntryFare()); Fares.Add(new StateTaxFare(taxiStarted)); Fares.Add(new NightSurchargeFare(taxiStarted)); Fares.Add(new PeakHourWeekdayFare(taxiStarted));   SetTotalFare(Fares); }   private void SetTotalFare(IEnumerable<IFare> fares) { TotalFare = (_totalFare ?? 0) + TaxiFareHelper.GetTotalFare(fares); } ....   } }   TotalViewModel subscribes to events, TaxiStartedEvent and rest. When TaxiStartedEvent gets invoked it calls the OnTaxiStarted method which sets the total fare which includes entry fee, state tax, nightly surcharge, peak hour weekday fare.   Note that TotalViewModel derives from ObservableBase which implements the method RaisePropertyChanged which we are invoking in Set of TotalFare property, i.e, once we update the TotalFare property it raises an the event that  allows the TotalFare text box to fetch the new value through the data context. ViewModel is communicating with View through data context and it has no knowledge about View, helping in loose coupling of ViewModel and View.   I have attached the source code (.Net 4.0, Prism 4.0, VS 2010) , download and play with it and don’t forget to leave your comments.  

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  • Red Gate Coder interviews: Alex Davies

    - by Michael Williamson
    Alex Davies has been a software engineer at Red Gate since graduating from university, and is currently busy working on .NET Demon. We talked about tackling parallel programming with his actors framework, a scientific approach to debugging, and how JavaScript is going to affect the programming languages we use in years to come. So, if we start at the start, how did you get started in programming? When I was seven or eight, I was given a BBC Micro for Christmas. I had asked for a Game Boy, but my dad thought it would be better to give me a proper computer. For a year or so, I only played games on it, but then I found the user guide for writing programs in it. I gradually started doing more stuff on it and found it fun. I liked creating. As I went into senior school I continued to write stuff on there, trying to write games that weren’t very good. I got a real computer when I was fourteen and found ways to write BASIC on it. Visual Basic to start with, and then something more interesting than that. How did you learn to program? Was there someone helping you out? Absolutely not! I learnt out of a book, or by experimenting. I remember the first time I found a loop, I was like “Oh my God! I don’t have to write out the same line over and over and over again any more. It’s amazing!” When did you think this might be something that you actually wanted to do as a career? For a long time, I thought it wasn’t something that you would do as a career, because it was too much fun to be a career. I thought I’d do chemistry at university and some kind of career based on chemical engineering. And then I went to a careers fair at school when I was seventeen or eighteen, and it just didn’t interest me whatsoever. I thought “I could be a programmer, and there’s loads of money there, and I’m good at it, and it’s fun”, but also that I shouldn’t spoil my hobby. Now I don’t really program in my spare time any more, which is a bit of a shame, but I program all the rest of the time, so I can live with it. Do you think you learnt much about programming at university? Yes, definitely! I went into university knowing how to make computers do anything I wanted them to do. However, I didn’t have the language to talk about algorithms, so the algorithms course in my first year was massively important. Learning other language paradigms like functional programming was really good for breadth of understanding. Functional programming influences normal programming through design rather than actually using it all the time. I draw inspiration from it to write imperative programs which I think is actually becoming really fashionable now, but I’ve been doing it for ages. I did it first! There were also some courses on really odd programming languages, a bit of Prolog, a little bit of C. Having a little bit of each of those is something that I would have never done on my own, so it was important. And then there are knowledge-based courses which are about not programming itself but things that have been programmed like TCP. Those are really important for examples for how to approach things. Did you do any internships while you were at university? Yeah, I spent both of my summers at the same company. I thought I could code well before I went there. Looking back at the crap that I produced, it was only surpassed in its crappiness by all of the other code already in that company. I’m so much better at writing nice code now than I used to be back then. Was there just not a culture of looking after your code? There was, they just didn’t hire people for their abilities in that area. They hired people for raw IQ. The first indicator of it going wrong was that they didn’t have any computer scientists, which is a bit odd in a programming company. But even beyond that they didn’t have people who learnt architecture from anyone else. Most of them had started straight out of university, so never really had experience or mentors to learn from. There wasn’t the experience to draw from to teach each other. In the second half of my second internship, I was being given tasks like looking at new technologies and teaching people stuff. Interns shouldn’t be teaching people how to do their jobs! All interns are going to have little nuggets of things that you don’t know about, but they shouldn’t consistently be the ones who know the most. It’s not a good environment to learn. I was going to ask how you found working with people who were more experienced than you… When I reached Red Gate, I found some people who were more experienced programmers than me, and that was difficult. I’ve been coding since I was tiny. At university there were people who were cleverer than me, but there weren’t very many who were more experienced programmers than me. During my internship, I didn’t find anyone who I classed as being a noticeably more experienced programmer than me. So, it was a shock to the system to have valid criticisms rather than just formatting criticisms. However, Red Gate’s not so big on the actual code review, at least it wasn’t when I started. We did an entire product release and then somebody looked over all of the UI of that product which I’d written and say what they didn’t like. By that point, it was way too late and I’d disagree with them. Do you think the lack of code reviews was a bad thing? I think if there’s going to be any oversight of new people, then it should be continuous rather than chunky. For me I don’t mind too much, I could go out and get oversight if I wanted it, and in those situations I felt comfortable without it. If I was managing the new person, then maybe I’d be keener on oversight and then the right way to do it is continuously and in very, very small chunks. Have you had any significant projects you’ve worked on outside of a job? When I was a teenager I wrote all sorts of stuff. I used to write games, I derived how to do isomorphic projections myself once. I didn’t know what the word was so I couldn’t Google for it, so I worked it out myself. It was horrifically complicated. But it sort of tailed off when I started at university, and is now basically zero. If I do side-projects now, they tend to be work-related side projects like my actors framework, NAct, which I started in a down tools week. Could you explain a little more about NAct? It is a little C# framework for writing parallel code more easily. Parallel programming is difficult when you need to write to shared data. Sometimes parallel programming is easy because you don’t need to write to shared data. When you do need to access shared data, you could just have your threads pile in and do their work, but then you would screw up the data because the threads would trample on each other’s toes. You could lock, but locks are really dangerous if you’re using more than one of them. You get interactions like deadlocks, and that’s just nasty. Actors instead allows you to say this piece of data belongs to this thread of execution, and nobody else can read it. If you want to read it, then ask that thread of execution for a piece of it by sending a message, and it will send the data back by a message. And that avoids deadlocks as long as you follow some obvious rules about not making your actors sit around waiting for other actors to do something. There are lots of ways to write actors, NAct allows you to do it as if it was method calls on other objects, which means you get all the strong type-safety that C# programmers like. Do you think that this is suitable for the majority of parallel programming, or do you think it’s only suitable for specific cases? It’s suitable for most difficult parallel programming. If you’ve just got a hundred web requests which are all independent of each other, then I wouldn’t bother because it’s easier to just spin them up in separate threads and they can proceed independently of each other. But where you’ve got difficult parallel programming, where you’ve got multiple threads accessing multiple bits of data in multiple ways at different times, then actors is at least as good as all other ways, and is, I reckon, easier to think about. When you’re using actors, you presumably still have to write your code in a different way from you would otherwise using single-threaded code. You can’t use actors with any methods that have return types, because you’re not allowed to call into another actor and wait for it. If you want to get a piece of data out of another actor, then you’ve got to use tasks so that you can use “async” and “await” to await asynchronously for it. But other than that, you can still stick things in classes so it’s not too different really. Rather than having thousands of objects with mutable state, you can use component-orientated design, where there are only a few mutable classes which each have a small number of instances. Then there can be thousands of immutable objects. If you tend to do that anyway, then actors isn’t much of a jump. If I’ve already built my system without any parallelism, how hard is it to add actors to exploit all eight cores on my desktop? Usually pretty easy. If you can identify even one boundary where things look like messages and you have components where some objects live on one side and these other objects live on the other side, then you can have a granddaddy object on one side be an actor and it will parallelise as it goes across that boundary. Not too difficult. If we do get 1000-core desktop PCs, do you think actors will scale up? It’s hard. There are always in the order of twenty to fifty actors in my whole program because I tend to write each component as actors, and I tend to have one instance of each component. So this won’t scale to a thousand cores. What you can do is write data structures out of actors. I use dictionaries all over the place, and if you need a dictionary that is going to be accessed concurrently, then you could build one of those out of actors in no time. You can use queuing to marshal requests between different slices of the dictionary which are living on different threads. So it’s like a distributed hash table but all of the chunks of it are on the same machine. That means that each of these thousand processors has cached one small piece of the dictionary. I reckon it wouldn’t be too big a leap to start doing proper parallelism. Do you think it helps if actors get baked into the language, similarly to Erlang? Erlang is excellent in that it has thread-local garbage collection. C# doesn’t, so there’s a limit to how well C# actors can possibly scale because there’s a single garbage collected heap shared between all of them. When you do a global garbage collection, you’ve got to stop all of the actors, which is seriously expensive, whereas in Erlang garbage collections happen per-actor, so they’re insanely cheap. However, Erlang deviated from all the sensible language design that people have used recently and has just come up with crazy stuff. You can definitely retrofit thread-local garbage collection to .NET, and then it’s quite well-suited to support actors, even if it’s not baked into the language. Speaking of language design, do you have a favourite programming language? I’ll choose a language which I’ve never written before. I like the idea of Scala. It sounds like C#, only with some of the niggles gone. I enjoy writing static types. It means you don’t have to writing tests so much. When you say it doesn’t have some of the niggles? C# doesn’t allow the use of a property as a method group. It doesn’t have Scala case classes, or sum types, where you can do a switch statement and the compiler checks that you’ve checked all the cases, which is really useful in functional-style programming. Pattern-matching, in other words. That’s actually the major niggle. C# is pretty good, and I’m quite happy with C#. And what about going even further with the type system to remove the need for tests to something like Haskell? Or is that a step too far? I’m quite a pragmatist, I don’t think I could deal with trying to write big systems in languages with too few other users, especially when learning how to structure things. I just don’t know anyone who can teach me, and the Internet won’t teach me. That’s the main reason I wouldn’t use it. If I turned up at a company that writes big systems in Haskell, I would have no objection to that, but I wouldn’t instigate it. What about things in C#? For instance, there’s contracts in C#, so you can try to statically verify a bit more about your code. Do you think that’s useful, or just not worthwhile? I’ve not really tried it. My hunch is that it needs to be built into the language and be quite mathematical for it to work in real life, and that doesn’t seem to have ended up true for C# contracts. I don’t think anyone who’s tried them thinks they’re any good. I might be wrong. On a slightly different note, how do you like to debug code? I think I’m quite an odd debugger. I use guesswork extremely rarely, especially if something seems quite difficult to debug. I’ve been bitten spending hours and hours on guesswork and not being scientific about debugging in the past, so now I’m scientific to a fault. What I want is to see the bug happening in the debugger, to step through the bug happening. To watch the program going from a valid state to an invalid state. When there’s a bug and I can’t work out why it’s happening, I try to find some piece of evidence which places the bug in one section of the code. From that experiment, I binary chop on the possible causes of the bug. I suppose that means binary chopping on places in the code, or binary chopping on a stage through a processing cycle. Basically, I’m very stupid about how I debug. I won’t make any guesses, I won’t use any intuition, I will only identify the experiment that’s going to binary chop most effectively and repeat rather than trying to guess anything. I suppose it’s quite top-down. Is most of the time then spent in the debugger? Absolutely, if at all possible I will never debug using print statements or logs. I don’t really hold much stock in outputting logs. If there’s any bug which can be reproduced locally, I’d rather do it in the debugger than outputting logs. And with SmartAssembly error reporting, there’s not a lot that can’t be either observed in an error report and just fixed, or reproduced locally. And in those other situations, maybe I’ll use logs. But I hate using logs. You stare at the log, trying to guess what’s going on, and that’s exactly what I don’t like doing. You have to just look at it and see does this look right or wrong. We’ve covered how you get to grip with bugs. How do you get to grips with an entire codebase? I watch it in the debugger. I find little bugs and then try to fix them, and mostly do it by watching them in the debugger and gradually getting an understanding of how the code works using my process of binary chopping. I have to do a lot of reading and watching code to choose where my slicing-in-half experiment is going to be. The last time I did it was SmartAssembly. The old code was a complete mess, but at least it did things top to bottom. There wasn’t too much of some of the big abstractions where flow of control goes all over the place, into a base class and back again. Code’s really hard to understand when that happens. So I like to choose a little bug and try to fix it, and choose a bigger bug and try to fix it. Definitely learn by doing. I want to always have an aim so that I get a little achievement after every few hours of debugging. Once I’ve learnt the codebase I might be able to fix all the bugs in an hour, but I’d rather be using them as an aim while I’m learning the codebase. If I was a maintainer of a codebase, what should I do to make it as easy as possible for you to understand? Keep distinct concepts in different places. And name your stuff so that it’s obvious which concepts live there. You shouldn’t have some variable that gets set miles up the top of somewhere, and then is read miles down to choose some later behaviour. I’m talking from a very much SmartAssembly point of view because the old SmartAssembly codebase had tons and tons of these things, where it would read some property of the code and then deal with it later. Just thousands of variables in scope. Loads of things to think about. If you can keep concepts separate, then it aids me in my process of fixing bugs one at a time, because each bug is going to more or less be understandable in the one place where it is. And what about tests? Do you think they help at all? I’ve never had the opportunity to learn a codebase which has had tests, I don’t know what it’s like! What about when you’re actually developing? How useful do you find tests in finding bugs or regressions? Finding regressions, absolutely. Running bits of code that would be quite hard to run otherwise, definitely. It doesn’t happen very often that a test finds a bug in the first place. I don’t really buy nebulous promises like tests being a good way to think about the spec of the code. My thinking goes something like “This code works at the moment, great, ship it! Ah, there’s a way that this code doesn’t work. Okay, write a test, demonstrate that it doesn’t work, fix it, use the test to demonstrate that it’s now fixed, and keep the test for future regressions.” The most valuable tests are for bugs that have actually happened at some point, because bugs that have actually happened at some point, despite the fact that you think you’ve fixed them, are way more likely to appear again than new bugs are. Does that mean that when you write your code the first time, there are no tests? Often. The chance of there being a bug in a new feature is relatively unaffected by whether I’ve written a test for that new feature because I’m not good enough at writing tests to think of bugs that I would have written into the code. So not writing regression tests for all of your code hasn’t affected you too badly? There are different kinds of features. Some of them just always work, and are just not flaky, they just continue working whatever you throw at them. Maybe because the type-checker is particularly effective around them. Writing tests for those features which just tend to always work is a waste of time. And because it’s a waste of time I’ll tend to wait until a feature has demonstrated its flakiness by having bugs in it before I start trying to test it. You can get a feel for whether it’s going to be flaky code as you’re writing it. I try to write it to make it not flaky, but there are some things that are just inherently flaky. And very occasionally, I’ll think “this is going to be flaky” as I’m writing, and then maybe do a test, but not most of the time. How do you think your programming style has changed over time? I’ve got clearer about what the right way of doing things is. I used to flip-flop a lot between different ideas. Five years ago I came up with some really good ideas and some really terrible ideas. All of them seemed great when I thought of them, but they were quite diverse ideas, whereas now I have a smaller set of reliable ideas that are actually good for structuring code. So my code is probably more similar to itself than it used to be back in the day, when I was trying stuff out. I’ve got more disciplined about encapsulation, I think. There are operational things like I use actors more now than I used to, and that forces me to use immutability more than I used to. The first code that I wrote in Red Gate was the memory profiler UI, and that was an actor, I just didn’t know the name of it at the time. I don’t really use object-orientation. By object-orientation, I mean having n objects of the same type which are mutable. I want a constant number of objects that are mutable, and they should be different types. I stick stuff in dictionaries and then have one thing that owns the dictionary and puts stuff in and out of it. That’s definitely a pattern that I’ve seen recently. I think maybe I’m doing functional programming. Possibly. It’s plausible. If you had to summarise the essence of programming in a pithy sentence, how would you do it? Programming is the form of art that, without losing any of the beauty of architecture or fine art, allows you to produce things that people love and you make money from. So you think it’s an art rather than a science? It’s a little bit of engineering, a smidgeon of maths, but it’s not science. Like architecture, programming is on that boundary between art and engineering. If you want to do it really nicely, it’s mostly art. You can get away with doing architecture and programming entirely by having a good engineering mind, but you’re not going to produce anything nice. You’re not going to have joy doing it if you’re an engineering mind. Architects who are just engineering minds are not going to enjoy their job. I suppose engineering is the foundation on which you build the art. Exactly. How do you think programming is going to change over the next ten years? There will be an unfortunate shift towards dynamically-typed languages, because of JavaScript. JavaScript has an unfair advantage. JavaScript’s unfair advantage will cause more people to be exposed to dynamically-typed languages, which means other dynamically-typed languages crop up and the best features go into dynamically-typed languages. Then people conflate the good features with the fact that it’s dynamically-typed, and more investment goes into dynamically-typed languages. They end up better, so people use them. What about the idea of compiling other languages, possibly statically-typed, to JavaScript? It’s a reasonable idea. I would like to do it, but I don’t think enough people in the world are going to do it to make it pick up. The hordes of beginners are the lifeblood of a language community. They are what makes there be good tools and what makes there be vibrant community websites. And any particular thing which is the same as JavaScript only with extra stuff added to it, although it might be technically great, is not going to have the hordes of beginners. JavaScript is always to be quickest and easiest way for a beginner to start programming in the browser. And dynamically-typed languages are great for beginners. Compilers are pretty scary and beginners don’t write big code. And having your errors come up in the same place, whether they’re statically checkable errors or not, is quite nice for a beginner. If someone asked me to teach them some programming, I’d teach them JavaScript. If dynamically-typed languages are great for beginners, when do you think the benefits of static typing start to kick in? The value of having a statically typed program is in the tools that rely on the static types to produce a smooth IDE experience rather than actually telling me my compile errors. And only once you’re experienced enough a programmer that having a really smooth IDE experience makes a blind bit of difference, does static typing make a blind bit of difference. So it’s not really about size of codebase. If I go and write up a tiny program, I’m still going to get value out of writing it in C# using ReSharper because I’m experienced with C# and ReSharper enough to be able to write code five times faster if I have that help. Any other visions of the future? Nobody’s going to use actors. Because everyone’s going to be running on single-core VMs connected over network-ready protocols like JSON over HTTP. So, parallelism within one operating system is going to die. But until then, you should use actors. More Red Gater Coder interviews

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  • Why should you choose Oracle WebLogic 12c instead of JBoss EAP 6?

    - by Ricardo Ferreira
    In this post, I will cover some technical differences between Oracle WebLogic 12c and JBoss EAP 6, which was released a couple days ago from Red Hat. This article claims to help you in the evaluation of key points that you should consider when choosing for an Java EE application server. In the following sections, I will present to you some important aspects that most customers ask us when they are seriously evaluating for an middleware infrastructure, specially if you are considering JBoss for some reason. I would suggest that you keep the following question in mind while you are reading the points: "Why should I choose JBoss instead of WebLogic?" 1) Multi Datacenter Deployment and Clustering - D/R ("Disaster & Recovery") architecture support is embedded on the WebLogic Server 12c product. JBoss EAP 6 on the other hand has no direct D/R support included, Red Hat relies on third-part tools with higher prices. When you consider a middleware solution to host your business critical application, you should worry with every architectural aspect that are related with the solution. Fail-over support is one little aspect of a truly reliable solution. If you do not worry about D/R, your solution will not be reliable. Having said that, with Red Hat and JBoss EAP 6, you have this extra cost that will increase considerably the total cost of ownership of the solution. As we commonly hear from analysts, open-source are not so cheaper when you start seeing the big picture. - WebLogic Server 12c supports advanced LAN clustering, detection of death servers and have a common alert framework. JBoss EAP 6 on the other hand has limited LAN clustering support with no server death detection. They do not generate any alerts when servers goes down (only if you buy JBoss ON which is a separated technology, but until now does not support JBoss EAP 6) and manual intervention are required when servers goes down. In most cases, admin people must rely on "kill -9", "tail -f someFile.log" and "ps ax | grep java" commands to manage failures and clustering anomalies. - WebLogic Server 12c supports the concept of Node Manager, which is a separated process that runs on the physical | virtual servers that allows extend the administration of the cluster to WebLogic managed servers that are often distributed across multiple machines and geographic locations. JBoss EAP 6 on the other hand has no equivalent technology. Whole server instances must be managed individually. - WebLogic Server 12c Node Manager supports Coherence to boost performance when managing servers. JBoss EAP 6 on the other hand has no similar technology. There is no way to coordinate JBoss and infiniband instances provided by JBoss using high throughput and low latency protocols like InfiniBand. The Node Manager feature also allows another very important feature that JBoss EAP lacks: secure the administration. When using WebLogic Node Manager, all the administration tasks are sent to the managed servers in a secure tunel protected by a certificate, which means that the transport layer that separates the WebLogic administration console from the managed servers are secured by SSL. - WebLogic Server 12c are now integrated with OTD ("Oracle Traffic Director") which is a web server technology derived from the former Sun iPlanet Web Server. This software complements the web server support offered by OHS ("Oracle HTTP Server"). Using OTD, WebLogic instances are load-balanced by a high powerful software that knows how to handle SDP ("Socket Direct Protocol") over InfiniBand, which boost performance when used with engineered systems technologies like Oracle Exalogic Elastic Cloud. JBoss EAP 6 on the other hand only offers support to Apache Web Server with custom modules created to deal with JBoss clusters, but only across standard TCP/IP networks.  2) Application and Runtime Diagnostics - WebLogic Server 12c have diagnostics capabilities embedded on the server called WLDF ("WebLogic Diagnostic Framework") so there is no need to rely on third-part tools. JBoss EAP 6 on the other hand has no diagnostics capabilities. Their only diagnostics tool is the log generated by the application server. Admin people are encouraged to analyse thousands of log lines to find out what is going on. - WebLogic Server 12c complement WLDF with JRockit MC ("Mission Control"), which provides to administrators and developers a complete insight about the JVM performance, behavior and possible bottlenecks. WebLogic Server 12c also have an classloader analysis tool embedded, and even a log analyzer tool that enables administrators and developers to view logs of multiple servers at the same time. JBoss EAP 6 on the other hand relies on third-part tools to do something similar. Again, only log searching are offered to find out whats going on. - WebLogic Server 12c offers end-to-end traceability and monitoring available through Oracle EM ("Enterprise Manager"), including monitoring of business transactions that flows through web servers, ESBs, application servers and database servers, all of this with high deep JVM analysis and diagnostics. JBoss EAP 6 on the other hand, even using JBoss ON ("Operations Network"), which is a separated technology, does not support those features. Red Hat relies on third-part tools to provide direct Oracle database traceability across JVMs. One of those tools are Oracle EM for non-Oracle middleware that manage JBoss, Tomcat, Websphere and IIS transparently. - WebLogic Server 12c with their JRockit support offers a tool called JRockit Flight Recorder, which can give developers a complete visibility of a certain period of application production monitoring with zero extra overhead. This automatic recording allows you to deep analyse threads latency, memory leaks, thread contention, resource utilization, stack overflow damages and GC ("Garbage Collection") cycles, to observe in real time stop-the-world phenomenons, generational, reference count and parallel collects and mutator threads analysis. JBoss EAP 6 don't even dream to support something similar, even because they don't have their own JVM. 3) Application Server Administration - WebLogic Server 12c offers a complete administration console complemented with scripting and macro-like recording capabilities. A single WebLogic console can managed up to hundreds of WebLogic servers belonging to the same domain. JBoss EAP 6 on the other hand has a limited console and provides a XML centric administration. JBoss, after ten years, started the development of a rudimentary centralized administration that still leave a lot of administration tasks aside, so admin people and developers must touch scripts and XML configuration files for most advanced and even simple administration tasks. This lead applications to error prone and risky deployments. Even using JBoss ON, JBoss EAP are not able to offer decent administration features for admin people which must be high skilled in JBoss internal architecture and its managing capabilities. - Oracle EM is available to manage multiple domains, databases, application servers, operating systems and virtualization, with a complete end-to-end visibility. JBoss ON does not provide management capabilities across the complete architecture, only basic monitoring. Even deployment must be done aside JBoss ON which does no integrate well with others softwares than JBoss. Until now, JBoss ON does not supports JBoss EAP 6, so even their minimal support for JBoss are not available for JBoss EAP 6 leaving customers uncovered and subject to high skilled JBoss admin people. - WebLogic Server 12c has the same administration model whatever is the topology selected by the customer. JBoss EAP 6 on the other hand differentiates between two operational models: standalone-mode and domain-mode, that are not consistent with each other. Depending on the mode used, the administration skill is different. - WebLogic Server 12c has no point-of-failures processes, and it does not need to define any specialized server. Domain model in WebLogic is available for years (at least ten years or more) and is production proven. JBoss EAP 6 on the other hand needs special processes to garantee JBoss integrity, the PC ("Process-Controller") and the HC ("Host-Controller"). Different from WebLogic, the domain model in JBoss is quite new (one year at tops) of maturity, and need to mature considerably until start doing things like WebLogic domain model does. - WebLogic Server 12c supports parallel deployment model which enables some artifacts being deployed at the same time. JBoss EAP 6 on the other hand does not have any similar feature. Every deployment are done atomically in the containers. This means that if you have a huge EAR (an EAR of 120 MB of size for instance) and deploy onto JBoss EAP 6, this EAR will take some minutes in order to starting accept thread requests. The same EAR deployed onto WebLogic Server 12c will reduce the deployment time at least in 2X compared to JBoss. 4) Support and Upgrades - WebLogic Server 12c has patch management available. JBoss EAP 6 on the other hand has no patch management available, each JBoss EAP instance should be patched manually. To achieve such feature, you need to buy a separated technology called JBoss ON ("Operations Network") that manage this type of stuff. But until now, JBoss ON does not support JBoss EAP 6 so, in practice, JBoss EAP 6 does not have this feature. - WebLogic Server 12c supports previuous WebLogic domains without any reconfiguration since its kernel is robust and mature since its creation in 1995. JBoss EAP 6 on the other hand has a proven lack of supportability between JBoss AS 4, 5, 6 and 7. Different kernels and messaging engines were implemented in JBoss stack in the last five years reveling their incapacity to create a well architected and proven middleware technology. - WebLogic Server 12c has patch prescription based on customer configuration. JBoss EAP 6 on the other hand has no such capability. People need to create ticket supports and have their installations revised by Red Hat support guys to gain some patch prescription from them. - Oracle WebLogic Server independent of the version has 8 years of support of new patches and has lifetime release of existing patches beyond that. JBoss EAP 6 on the other hand provides patches for a specific application server version up to 5 years after the release date. JBoss EAP 4 and previous versions had only 4 years. A good question that Red Hat will argue to answer is: "what happens when you find issues after year 5"?  5) RAC ("Real Application Clusters") Support - WebLogic Server 12c ships with a specific JDBC driver to leverage Oracle RAC clustering capabilities (Fast-Application-Notification, Transaction Affinity, Fast-Connection-Failover, etc). Oracle JDBC thin driver are also available. JBoss EAP 6 on the other hand ships only the standard Oracle JDBC thin driver. Load balancing with Oracle RAC are not supported. Manual intervention in case of planned or unplanned RAC downtime are necessary. In JBoss EAP 6, situation does not reestablish automatically after downtime. - WebLogic Server 12c has a feature called Active GridLink for Oracle RAC which provides up to 3X performance on OLTP applications. This seamless integration between WebLogic and Oracle database enable more value added to critical business applications leveraging their investments in Oracle database technology and Oracle middleware. JBoss EAP 6 on the other hand has no performance gains at all, even when admin people implement some kind of connection-pooling tuning. - WebLogic Server 12c also supports transaction and web session affinity to the Oracle RAC, which provides aditional gains of performance. This is particularly interesting if you are creating a reliable solution that are distributed not only in an LAN cluster, but into a different data center. JBoss EAP 6 on the other hand has no such support. 6) Standards and Technology Support - WebLogic Server 12c is fully Java EE 6 compatible and production ready since december of 2011. JBoss EAP 6 on the other hand became fully compatible with Java EE 6 only in the community version after three months, and production ready only in a few days considering that this article was written in June of 2012. Red Hat says that they are the masters of innovation and technology proliferation, but compared with Oracle and even other proprietary vendors like IBM, they historically speaking are lazy to deliver the most newest technologies and standards adherence. - Oracle is the steward of Java, driving innovation into the platform from commercial and open-source vendors. Red Hat on the other hand does not have its own JVM and relies on third-part JVMs to complete their application server offer. 95% of Red Hat customers are using Oracle HotSpot as JVM, which means that without Oracle involvement, their support are limited exclusively to the application server layer and we all know that most problems are happens in the JVM layer. - WebLogic Server 12c supports natively JDK 7, which empower developers to explore the maximum of the Java platform productivity when writing code. This feature differentiate WebLogic from others application servers (except GlassFish that are also managed by Oracle) because the usage of JDK 7 introduce such remarkable productivity features like the "try-with-resources" enhancement, catching multiple exceptions with one try block, Strings in the switch statements, JVM improvements in terms of JDBC, I/O, networking, security, concurrency and of course, the most important feature of Java 7: native support for multiple non-Java languages. More features regarding JDK 7 can be found here. JBoss EAP 6 on the other hand does not support JDK 7 officially, they comment in their community version that "Java SE 7 can be used with JBoss 7" which does not gives you any guarantees of enterprise support for JDK 7. - Oracle WebLogic Server 12c supports integration with Spring framework allowing Spring applications to use WebLogic special transaction manager, exposing bean interfaces to WebLogic MBeans to take advantage of all WebLogic monitoring and administration advantages. JBoss EAP 6 on the other hand has no special integration with Spring. In fact, Red Hat offers a suspicious package called "JBoss Web Platform" that in theory supports Spring, but in practice this package does not offers any special integration. It is just a facility for Red Hat customers to have support from both JBoss and Spring technology using the same customer support. 7) Lightweight Development - Oracle WebLogic Server 12c and Oracle GlassFish are completely integrated and can share applications without any modifications. Starting with the 12c version, WebLogic now understands natively GlassFish deployment descriptors and specific configurations in order to offer you a truly and reliable migration path from a community Java EE application server to a enterprise middleware product like WebLogic. JBoss EAP 6 on the other hand has no support to natively reuse an existing (or still in development) application from JBoss AS community server. Users of JBoss suffer of critical issues during deployment time that includes: changing the libraries and dependencies of the application, patching the DTD or XSD deployment descriptors, refactoring of the application layers due classloading issues and anomalies, rebuilding of persistence, business and web layers due issues with "usage of the certified version of an certain dependency" or "frameworks that Red Hat potentially does not recommend" etc. If you have the culture or enterprise IT directive of developing Java EE applications using community middleware to in a certain future, transition to enterprise (supported by a vendor) middleware, Oracle WebLogic plus Oracle GlassFish offers you a more sustainable solution. - WebLogic Server 12c has a very light ZIP distribution (less than 165 MB). JBoss EAP 6 ZIP size is around 130 MB, together with JBoss ON you have more 100 MB resulting in a higher download footprint. This is particularly interesting if you plan to use automated setup of application server instances (for example, to rapidly setup a development or staging environment) using Maven or Hudson. - WebLogic Server 12c has a complete integration with Maven allowing developers to setup WebLogic domains with few commands. Tasks like downloading WebLogic, installation, domain creation, data sources deployment are completely integrated. JBoss EAP 6 on the other hand has a limited offer integration with those tools.  - WebLogic Server 12c has a startup mode called WLX that turns-off EJB, JMS and JCA containers leaving enabled only the web container with Java EE 6 web profile. JBoss EAP 6 on the other hand has no such feature, you need to disable manually the containers that you do not want to use. - WebLogic Server 12c supports fastswap, which enables you to change classes without redeployment. This is particularly interesting if you are developing patches for the application that is already deployed and you do not want to redeploy the entire application. This is the same behavior that most application servers offers to JSP pages, but with WebLogic Server 12c, you have the same feature for Java classes in general. JBoss EAP 6 on the other hand has no such support. Even JBoss EAP 5 does not support this until now. 8) JMS and Messaging - WebLogic Server 12c has a proven and high scalable JMS implementation since its initial release in 1995. JBoss EAP 6 on the other hand has a still immature technology called HornetQ, which was introduced in JBoss EAP 5 replacing everything that was implemented in the previous versions. Red Hat loves to introduce new technologies across JBoss versions, playing around with customers and their investments. And when they are asked about why they have changed the implementation and caused such a mess, their answer is always: "the previous implementation was inadequate and not aligned with the community strategy so we are creating a new a improved one". This Red Hat practice leads to uncomfortable investments that in a near future (sometimes less than a year) will be affected in someway. - WebLogic Server 12c has troubleshooting and monitoring features included on the WebLogic console and WLDF. JBoss EAP 6 on the other hand has no direct monitoring on the console, activity is reflected only on the logs, no debug logs available in case of JMS issues. - WebLogic Server 12c has extremely good performance and scalability. JBoss EAP 6 on the other hand has a JMS storage mechanism relying on Oracle database or MySQL. This means that if an issue in production happens and Red Hat affirms that an performance issue is happening due to database problems, they will not support you on the performance issue. They will orient you to call Oracle instead. - WebLogic Server 12c supports messaging enterprise features like SAF ("Store and Forward"), Distributed Queues/Topics and Foreign JMS providers support that leverage JMS implementations without compromise developer code making things completely transparent. JBoss EAP 6 on the other hand do not even dream to support such features. 9) Caching and Grid - Coherence, which is the leading and most mature data grid technology from Oracle, is available since early 2000 and was integrated with WebLogic in 2009. Coherence and WebLogic clusters can be both managed from WebLogic administrative console. Even Node Manager supports Coherence. JBoss on the other hand discontinued JBoss Cache, which was their caching implementation just like they did with the messaging implementation (JBossMQ) which was a issue for long term customers. JBoss EAP 6 ships InfiniSpan version 1.0 which is immature and lack a proven record of successful cases and reliability. - WebLogic Server 12c has a feature called ActiveCache which uses Coherence to, without any code changes, replicate HTTP sessions from both WebLogic and other application servers like JBoss, Tomcat, Websphere, GlassFish and even Microsoft IIS. JBoss EAP 6 on the other hand does have such support and even when they do in the future, they probably will support only their own application server. - Coherence can be used to manage both L1 and L2 cache levels, providing support to Oracle TopLink and others JPA compliant implementations, even Hibernate. JBoss EAP 6 and Infinispan on the other hand supports only Hibernate. And most important of all: Infinispan does not have any successful case of L1 or L2 caching level support using Hibernate, which lead us to reflect about its viability. 10) Performance - WebLogic Server 12c is certified with Oracle Exalogic Elastic Cloud and can run unchanged applications at this engineered system. This approach can benefit customers from Exalogic optimization's of both kernel and JVM layers to boost performance in terms of 10X for web, OLTP, JMS and grid applications. JBoss EAP 6 on the other hand has no investment on engineered systems: customers do not have the choice to deploy on a Java ultra fast system if their project becomes relevant and performance issues are detected. - WebLogic Server 12c maintains a performance gain across each new release: starting on WebLogic 5.1, the overall performance gain has been close to 4X, which close to a 20% gain release by release. JBoss on the other hand does not provide SPECJAppServer or SPECJEnterprise performance benchmarks. Their so called "performance gains" remains hidden in their customer environments, which lead us to think if it is true or not since we will never get access to those environments. - WebLogic Server 12c has industry performance benchmarks with submissions across platforms and configurations leading SPECJ. Oracle WebLogic leads SPECJAppServer performance in multiple categories, fitting all customer topologies like: dual-node, single-node, multi-node and multi-node with RAC. JBoss... again, does not provide any SPECJAppServer performance benchmarks. - WebLogic Server 12c has a feature called work manager which allows your application to embrace new performance levels based on critical resource utilization of the CPUs usage. Work managers prioritizes work and allocates threads based on an execution model that takes into account administrator-defined parameters and actual run-time performance and throughput. JBoss EAP 6 on the other hand has no compared feature and probably they never will. Not supporting such feature like work managers, JBoss EAP 6 forces admin people and specially developers to uncover performance gains in a intrusive way, rewriting the code and doing performance refactorings. 11) Professional Services Support - WebLogic Server 12c and any other technology sold by Oracle give customers the possibility of hire OCS ("Oracle Consulting Services") to manage critical scenarios, deployment assistance of new applications, high skilled consultancy of architecture, best practices and people allocation together with customer teams. All OCS services are available without any restrictions, having the customer bought software from Oracle or just starting their implementation before any acquisition. JBoss EAP 6 or Red Hat to be more specifically, only offers professional services if you buy subscriptions from them. If you are developing a new critical application for your business and need the help of Red Hat for a serious issue or architecture decision, they will probably say: "OK... I can help you but after you buy subscriptions from me". Red Hat also does not allows their professional services consultants to manage environments that uses community based software. They will probably force you to first buy a subscription, download their "enterprise" version and them, optionally hire their consultants. - Oracle provides you our university to educate your team into our technologies, including of course specialized trainings of WebLogic application server. At any time and location, you can hire Oracle to train your team so you get trustful knowledge according to your specific needs. Certifications for the products are also available if your technical people desire to differentiate themselves as professionals. Red Hat on the other hand have a limited pool of resources to train your team in their technologies. Basically they are selling training and certification for RHEL ("Red Hat Enterprise Linux") but if you demand more specialized training in JBoss middleware, they will probably connect you to some "certified" partner localized training since they are apparently discontinuing their education center, at least here in Brazil. They were not able to reproduce their success with RHEL education to their middleware division since they need first sell the subscriptions to after gives you specialized training. And again, they only offer you specialized training based on their enterprise version (EAP in the case of JBoss) which means that the courses will be a quite outdated. There are reports of developers that took official training's from Red Hat at this year (2012) and in a certain JBoss advanced course, Red Hat supposedly covered JBossMQ as the messaging subsystem, and even the printed material provided was based on JBossMQ since the training was created for JBoss EAP 4.3. 12) Encouraging Transparency without Ulterior Motives - WebLogic Server 12c like any other software from Oracle can be downloaded any time from anywhere, you should only possess an OTN ("Oracle Technology Network") credential and you can download any enterprise software how many times you want. And is not some kind of "trial" version. It is the official binaries that will be running for ever in your data center. Oracle does not encourages the usage of "specific versions" of our software. The binaries you buy from Oracle are the same binaries anyone in the world could download and use for testing and personal education. JBoss EAP 6 on the other hand are not available for download unless you buy a subscription and get access to the Red Hat enterprise repositories. If you need to test, learn or just start creating your application using Red Hat's middleware software, you should download it from the community website. You are not allowed to download the enterprise version that, according to Red Hat are more secure, reliable and robust. But no one of us want to start the development of a software with an unsecured, unreliable and not scalable middleware right? So what you do? You are "invited" by Red Hat to buy subscriptions from them to get access to the "cool" version of the software. - WebLogic Server 12c prices are publicly available in the Oracle website. If you want to know right now how much WebLogic will cost to your organization, just click here and get access to our price list. In the case of WebLogic, check out the "US Oracle Technology Commercial Price List". Oracle also encourages you to get in touch with a sales representative to discuss discounts that would make possible the investment into our technology. But you are not required to do this, only if you are interested in buying our technology or maybe you want to discuss some discount scenarios. JBoss EAP 6 on the other hand does not have its cost publicly available in Red Hat's website or in any other media, at least is not so easy to get such information. The only link you will possibly find in their website is a "Contact a Sales Representative" link. This is not a very good relationship between an customer and an vendor. This is not an example of transparency, mainly when the software are sold as open. In this situations, customers expects to see the software prices publicly available, so they can have the chance to decide, based on the existing features of the software, if the cost is fair or not. Conclusion Oracle WebLogic is the most mature, secure, reliable and scalable Java EE application server of the market, and have a proven record of success around the globe to prove it's majority. Don't lose the chance to discover today how WebLogic could fit your needs and sustain your global IT middleware strategy, no matter if your strategy are completely based on the Cloud or not.

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  • Using HTML 5 SessionState to save rendered Page Content

    - by Rick Strahl
    HTML 5 SessionState and LocalStorage are very useful and super easy to use to manage client side state. For building rich client side or SPA style applications it's a vital feature to be able to cache user data as well as HTML content in order to swap pages in and out of the browser's DOM. What might not be so obvious is that you can also use the sessionState and localStorage objects even in classic server rendered HTML applications to provide caching features between pages. These APIs have been around for a long time and are supported by most relatively modern browsers and even all the way back to IE8, so you can use them safely in your Web applications. SessionState and LocalStorage are easy The APIs that make up sessionState and localStorage are very simple. Both object feature the same API interface which  is a simple, string based key value store that has getItem, setItem, removeitem, clear and  key methods. The objects are also pseudo array objects and so can be iterated like an array with  a length property and you have array indexers to set and get values with. Basic usage  for storing and retrieval looks like this (using sessionStorage, but the syntax is the same for localStorage - just switch the objects):// set var lastAccess = new Date().getTime(); if (sessionStorage) sessionStorage.setItem("myapp_time", lastAccess.toString()); // retrieve in another page or on a refresh var time = null; if (sessionStorage) time = sessionStorage.getItem("myapp_time"); if (time) time = new Date(time * 1); else time = new Date(); sessionState stores data that is browser session specific and that has a liftetime of the active browser session or window. Shut down the browser or tab and the storage goes away. localStorage uses the same API interface, but the lifetime of the data is permanently stored in the browsers storage area until deleted via code or by clearing out browser cookies (not the cache). Both sessionStorage and localStorage space is limited. The spec is ambiguous about this - supposedly sessionStorage should allow for unlimited size, but it appears that most WebKit browsers support only 2.5mb for either object. This means you have to be careful what you store especially since other applications might be running on the same domain and also use the storage mechanisms. That said 2.5mb worth of character data is quite a bit and would go a long way. The easiest way to get a feel for how sessionState and localStorage work is to look at a simple example. You can go check out the following example online in Plunker: http://plnkr.co/edit/0ICotzkoPjHaWa70GlRZ?p=preview which looks like this: Plunker is an online HTML/JavaScript editor that lets you write and run Javascript code and similar to JsFiddle, but a bit cleaner to work in IMHO (thanks to John Papa for turning me on to it). The sample has two text boxes with counts that update session/local storage every time you click the related button. The counts are 'cached' in Session and Local storage. The point of these examples is that both counters survive full page reloads, and the LocalStorage counter survives a complete browser shutdown and restart. Go ahead and try it out by clicking the Reload button after updating both counters and then shutting down the browser completely and going back to the same URL (with the same browser). What you should see is that reloads leave both counters intact at the counted values, while a browser restart will leave only the local storage counter intact. The code to deal with the SessionStorage (and LocalStorage not shown here) in the example is isolated into a couple of wrapper methods to simplify the code: function getSessionCount() { var count = 0; if (sessionStorage) { var count = sessionStorage.getItem("ss_count"); count = !count ? 0 : count * 1; } $("#txtSession").val(count); return count; } function setSessionCount(count) { if (sessionStorage) sessionStorage.setItem("ss_count", count.toString()); } These two functions essentially load and store a session counter value. The two key methods used here are: sessionStorage.getItem(key); sessionStorage.setItem(key,stringVal); Note that the value given to setItem and return by getItem has to be a string. If you pass another type you get an error. Don't let that limit you though - you can easily enough store JSON data in a variable so it's quite possible to pass complex objects and store them into a single sessionStorage value:var user = { name: "Rick", id="ricks", level=8 } sessionStorage.setItem("app_user",JSON.stringify(user)); to retrieve it:var user = sessionStorage.getItem("app_user"); if (user) user = JSON.parse(user); Simple! If you're using the Chrome Developer Tools (F12) you can also check out the session and local storage state on the Resource tab:   You can also use this tool to refresh or remove entries from storage. What we just looked at is a purely client side implementation where a couple of counters are stored. For rich client centric AJAX applications sessionStorage and localStorage provide a very nice and simple API to store application state while the application is running. But you can also use these storage mechanisms to manage server centric HTML applications when you combine server rendering with some JavaScript to perform client side data caching. You can both store some state information and data on the client (ie. store a JSON object and carry it forth between server rendered HTML requests) or you can use it for good old HTTP based caching where some rendered HTML is saved and then restored later. Let's look at the latter with a real life example. Why do I need Client-side Page Caching for Server Rendered HTML? I don't know about you, but in a lot of my existing server driven applications I have lists that display a fair amount of data. Typically these lists contain links to then drill down into more specific data either for viewing or editing. You can then click on a link and go off to a detail page that provides more concise content. So far so good. But now you're done with the detail page and need to get back to the list, so you click on a 'bread crumbs trail' or an application level 'back to list' button and… …you end up back at the top of the list - the scroll position, the current selection in some cases even filters conditions - all gone with the wind. You've left behind the state of the list and are starting from scratch in your browsing of the list from the top. Not cool! Sound familiar? This a pretty common scenario with server rendered HTML content where it's so common to display lists to drill into, only to lose state in the process of returning back to the original list. Look at just about any traditional forums application, or even StackOverFlow to see what I mean here. Scroll down a bit to look at a post or entry, drill in then use the bread crumbs or tab to go back… In some cases returning to the top of a list is not a big deal. On StackOverFlow that sort of works because content is turning around so quickly you probably want to actually look at the top posts. Not always though - if you're browsing through a list of search topics you're interested in and drill in there's no way back to that position. Essentially anytime you're actively browsing the items in the list, that's when state becomes important and if it's not handled the user experience can be really disrupting. Content Caching If you're building client centric SPA style applications this is a fairly easy to solve problem - you tend to render the list once and then update the page content to overlay the detail content, only hiding the list temporarily until it's used again later. It's relatively easy to accomplish this simply by hiding content on the page and later making it visible again. But if you use server rendered content, hanging on to all the detail like filters, selections and scroll position is not quite as easy. Or is it??? This is where sessionStorage comes in handy. What if we just save the rendered content of a previous page, and then restore it when we return to this page based on a special flag that tells us to use the cached version? Let's see how we can do this. A real World Use Case Recently my local ISP asked me to help out with updating an ancient classifieds application. They had a very busy, local classifieds app that was originally an ASP classic application. The old app was - wait for it: frames based - and even though I lobbied against it, the decision was made to keep the frames based layout to allow rapid browsing of the hundreds of posts that are made on a daily basis. The primary reason they wanted this was precisely for the ability to quickly browse content item by item. While I personally hate working with Frames, I have to admit that the UI actually works well with the frames layout as long as you're running on a large desktop screen. You can check out the frames based desktop site here: http://classifieds.gorge.net/ However when I rebuilt the app I also added a secondary view that doesn't use frames. The main reason for this of course was for mobile displays which work horribly with frames. So there's a somewhat mobile friendly interface to the interface, which ditches the frames and uses some responsive design tweaking for mobile capable operation: http://classifeds.gorge.net/mobile  (or browse the base url with your browser width under 800px)   Here's what the mobile, non-frames view looks like:   As you can see this means that the list of classifieds posts now is a list and there's a separate page for drilling down into the item. And of course… originally we ran into that usability issue I mentioned earlier where the browse, view detail, go back to the list cycle resulted in lost list state. Originally in mobile mode you scrolled through the list, found an item to look at and drilled in to display the item detail. Then you clicked back to the list and BAM - you've lost your place. Because there are so many items added on a daily basis the full list is never fully loaded, but rather there's a "Load Additional Listings"  entry at the button. Not only did we originally lose our place when coming back to the list, but any 'additionally loaded' items are no longer there because the list was now rendering  as if it was the first page hit. The additional listings, and any filters, the selection of an item all were lost. Major Suckage! Using Client SessionStorage to cache Server Rendered Content To work around this problem I decided to cache the rendered page content from the list in SessionStorage. Anytime the list renders or is updated with Load Additional Listings, the page HTML is cached and stored in Session Storage. Any back links from the detail page or the login or write entry forms then point back to the list page with a back=true query string parameter. If the server side sees this parameter it doesn't render the part of the page that is cached. Instead the client side code retrieves the data from the sessionState cache and simply inserts it into the page. It sounds pretty simple, and the overall the process is really easy, but there are a few gotchas that I'll discuss in a minute. But first let's look at the implementation. Let's start with the server side here because that'll give a quick idea of the doc structure. As I mentioned the server renders data from an ASP.NET MVC view. On the list page when returning to the list page from the display page (or a host of other pages) looks like this: https://classifieds.gorge.net/list?back=True The query string value is a flag, that indicates whether the server should render the HTML. Here's what the top level MVC Razor view for the list page looks like:@model MessageListViewModel @{ ViewBag.Title = "Classified Listing"; bool isBack = !string.IsNullOrEmpty(Request.QueryString["back"]); } <form method="post" action="@Url.Action("list")"> <div id="SizingContainer"> @if (!isBack) { @Html.Partial("List_CommandBar_Partial", Model) <div id="PostItemContainer" class="scrollbox" xstyle="-webkit-overflow-scrolling: touch;"> @Html.Partial("List_Items_Partial", Model) @if (Model.RequireLoadEntry) { <div class="postitem loadpostitems" style="padding: 15px;"> <div id="LoadProgress" class="smallprogressright"></div> <div class="control-progress"> Load additional listings... </div> </div> } </div> } </div> </form> As you can see the query string triggers a conditional block that if set is simply not rendered. The content inside of #SizingContainer basically holds  the entire page's HTML sans the headers and scripts, but including the filter options and menu at the top. In this case this makes good sense - in other situations the fact that the menu or filter options might be dynamically updated might make you only cache the list rather than essentially the entire page. In this particular instance all of the content works and produces the proper result as both the list along with any filter conditions in the form inputs are restored. Ok, let's move on to the client. On the client there are two page level functions that deal with saving and restoring state. Like the counter example I showed earlier, I like to wrap the logic to save and restore values from sessionState into a separate function because they are almost always used in several places.page.saveData = function(id) { if (!sessionStorage) return; var data = { id: id, scroll: $("#PostItemContainer").scrollTop(), html: $("#SizingContainer").html() }; sessionStorage.setItem("list_html",JSON.stringify(data)); }; page.restoreData = function() { if (!sessionStorage) return; var data = sessionStorage.getItem("list_html"); if (!data) return null; return JSON.parse(data); }; The data that is saved is an object which contains an ID which is the selected element when the user clicks and a scroll position. These two values are used to reset the scroll position when the data is used from the cache. Finally the html from the #SizingContainer element is stored, which makes for the bulk of the document's HTML. In this application the HTML captured could be a substantial bit of data. If you recall, I mentioned that the server side code renders a small chunk of data initially and then gets more data if the user reads through the first 50 or so items. The rest of the items retrieved can be rather sizable. Other than the JSON deserialization that's Ok. Since I'm using SessionStorage the storage space has no immediate limits. Next is the core logic to handle saving and restoring the page state. At first though this would seem pretty simple, and in some cases it might be, but as the following code demonstrates there are a few gotchas to watch out for. Here's the relevant code I use to save and restore:$( function() { … var isBack = getUrlEncodedKey("back", location.href); if (isBack) { // remove the back key from URL setUrlEncodedKey("back", "", location.href); var data = page.restoreData(); // restore from sessionState if (!data) { // no data - force redisplay of the server side default list window.location = "list"; return; } $("#SizingContainer").html(data.html); var el = $(".postitem[data-id=" + data.id + "]"); $(".postitem").removeClass("highlight"); el.addClass("highlight"); $("#PostItemContainer").scrollTop(data.scroll); setTimeout(function() { el.removeClass("highlight"); }, 2500); } else if (window.noFrames) page.saveData(null); // save when page loads $("#SizingContainer").on("click", ".postitem", function() { var id = $(this).attr("data-id"); if (!id) return true; if (window.noFrames) page.saveData(id); var contentFrame = window.parent.frames["Content"]; if (contentFrame) contentFrame.location.href = "show/" + id; else window.location.href = "show/" + id; return false; }); … The code starts out by checking for the back query string flag which triggers restoring from the client cache. If cached the cached data structure is read from sessionStorage. It's important here to check if data was returned. If the user had back=true on the querystring but there is no cached data, he likely bookmarked this page or otherwise shut down the browser and came back to this URL. In that case the server didn't render any detail and we have no cached data, so all we can do is redirect to the original default list view using window.location. If we continued the page would render no data - so make sure to always check the cache retrieval result. Always! If there is data the it's loaded and the data.html data is restored back into the document by simply injecting the HTML back into the document's #SizingContainer element:$("#SizingContainer").html(data.html); It's that simple and it's quite quick even with a fully loaded list of additional items and on a phone. The actual HTML data is stored to the cache on every page load initially and then again when the user clicks on an element to navigate to a particular listing. The former ensures that the client cache always has something in it, and the latter updates with additional information for the selected element. For the click handling I use a data-id attribute on the list item (.postitem) in the list and retrieve the id from that. That id is then used to navigate to the actual entry as well as storing that Id value in the saved cached data. The id is used to reset the selection by searching for the data-id value in the restored elements. The overall process of this save/restore process is pretty straight forward and it doesn't require a bunch of code, yet it yields a huge improvement in the usability of the site on mobile devices (or anybody who uses the non-frames view). Some things to watch out for As easy as it conceptually seems to simply store and retrieve cached content, you have to be quite aware what type of content you are caching. The code above is all that's specific to cache/restore cycle and it works, but it took a few tweaks to the rest of the script code and server code to make it all work. There were a few gotchas that weren't immediately obvious. Here are a few things to pay attention to: Event Handling Logic Timing of manipulating DOM events Inline Script Code Bookmarking to the Cache Url when no cache exists Do you have inline script code in your HTML? That script code isn't going to run if you restore from cache and simply assign or it may not run at the time you think it would normally in the DOM rendering cycle. JavaScript Event Hookups The biggest issue I ran into with this approach almost immediately is that originally I had various static event handlers hooked up to various UI elements that are now cached. If you have an event handler like:$("#btnSearch").click( function() {…}); that works fine when the page loads with server rendered HTML, but that code breaks when you now load the HTML from cache. Why? Because the elements you're trying to hook those events to may not actually be there - yet. Luckily there's an easy workaround for this by using deferred events. With jQuery you can use the .on() event handler instead:$("#SelectionContainer").on("click","#btnSearch", function() {…}); which monitors a parent element for the events and checks for the inner selector elements to handle events on. This effectively defers to runtime event binding, so as more items are added to the document bindings still work. For any cached content use deferred events. Timing of manipulating DOM Elements Along the same lines make sure that your DOM manipulation code follows the code that loads the cached content into the page so that you don't manipulate DOM elements that don't exist just yet. Ideally you'll want to check for the condition to restore cached content towards the top of your script code, but that can be tricky if you have components or other logic that might not all run in a straight line. Inline Script Code Here's another small problem I ran into: I use a DateTime Picker widget I built a while back that relies on the jQuery date time picker. I also created a helper function that allows keyboard date navigation into it that uses JavaScript logic. Because MVC's limited 'object model' the only way to embed widget content into the page is through inline script. This code broken when I inserted the cached HTML into the page because the script code was not available when the component actually got injected into the page. As the last bullet - it's a matter of timing. There's no good work around for this - in my case I pulled out the jQuery date picker and relied on native <input type="date" /> logic instead - a better choice these days anyway, especially since this view is meant to be primarily to serve mobile devices which actually support date input through the browser (unlike desktop browsers of which only WebKit seems to support it). Bookmarking Cached Urls When you cache HTML content you have to make a decision whether you cache on the client and also not render that same content on the server. In the Classifieds app I didn't render server side content so if the user comes to the page with back=True and there is no cached content I have to a have a Plan B. Typically this happens when somebody ends up bookmarking the back URL. The easiest and safest solution for this scenario is to ALWAYS check the cache result to make sure it exists and if not have a safe URL to go back to - in this case to the plain uncached list URL which amounts to effectively redirecting. This seems really obvious in hindsight, but it's easy to overlook and not see a problem until much later, when it's not obvious at all why the page is not rendering anything. Don't use <body> to replace Content Since we're practically replacing all the HTML in the page it may seem tempting to simply replace the HTML content of the <body> tag. Don't. The body tag usually contains key things that should stay in the page and be there when it loads. Specifically script tags and elements and possibly other embedded content. It's best to create a top level DOM element specifically as a placeholder container for your cached content and wrap just around the actual content you want to replace. In the app above the #SizingContainer is that container. Other Approaches The approach I've used for this application is kind of specific to the existing server rendered application we're running and so it's just one approach you can take with caching. However for server rendered content caching this is a pattern I've used in a few apps to retrofit some client caching into list displays. In this application I took the path of least resistance to the existing server rendering logic. Here are a few other ways that come to mind: Using Partial HTML Rendering via AJAXInstead of rendering the page initially on the server, the page would load empty and the client would render the UI by retrieving the respective HTML and embedding it into the page from a Partial View. This effectively makes the initial rendering and the cached rendering logic identical and removes the server having to decide whether this request needs to be rendered or not (ie. not checking for a back=true switch). All the logic related to caching is made on the client in this case. Using JSON Data and Client RenderingThe hardcore client option is to do the whole UI SPA style and pull data from the server and then use client rendering or databinding to pull the data down and render using templates or client side databinding with knockout/angular et al. As with the Partial Rendering approach the advantage is that there's no difference in the logic between pulling the data from cache or rendering from scratch other than the initial check for the cache request. Of course if the app is a  full on SPA app, then caching may not be required even - the list could just stay in memory and be hidden and reactivated. I'm sure there are a number of other ways this can be handled as well especially using  AJAX. AJAX rendering might simplify the logic, but it also complicates search engine optimization since there's no content loaded initially. So there are always tradeoffs and it's important to look at all angles before deciding on any sort of caching solution in general. State of the Session SessionState and LocalStorage are easy to use in client code and can be integrated even with server centric applications to provide nice caching features of content and data. In this post I've shown a very specific scenario of storing HTML content for the purpose of remembering list view data and state and making the browsing experience for lists a bit more friendly, especially if there's dynamically loaded content involved. If you haven't played with sessionStorage or localStorage I encourage you to give it a try. There's a lot of cool stuff that you can do with this beyond the specific scenario I've covered here… Resources Overview of localStorage (also applies to sessionStorage) Web Storage Compatibility Modernizr Test Suite© Rick Strahl, West Wind Technologies, 2005-2013Posted in JavaScript  HTML5  ASP.NET  MVC   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • A first look at ConfORM - Part 1

    - by thangchung
    All source codes for this post can be found at here.Have you ever heard of ConfORM is not? I have read it three months ago when I wrote an post about NHibernate and Autofac. At that time, this project really has just started and still in beta version, so I still do not really care much. But recently when reading a book by Jason Dentler NHibernate 3.0 Cookbook, I started to pay attention to it. Author have mentioned quite a lot of OSS in his book. And now again I have reviewed ConfORM once again. I have been involved in ConfORM development group on google and read some articles about it. Fabio Maulo spent a lot of work for the OSS, and I hope it will adapt a great way for NHibernate (because he contributed to NHibernate that). So what is ConfORM? It is stand for Configuration ORM, and it was trying to use a lot of heuristic model for identifying entities from C# code. Today, it's mostly Model First Driven development, so the first thing is to build the entity model. This is really important and we can see it is the heart of business software. Then we have to tell DB about the entity of this model. We often will use Inversion Engineering here, Database Schema is will create based on recently Entity Model. From now we will absolutely not interested in the DB again, only focus on the Entity Model.Fluent NHibenate really good, I liked this OSS. Sharp Architecture and has done so well in Fluent NHibernate integration with applications. A Multiple Database technical in Sharp Architecture is truly awesome. It can receive configuration, a connection string and a dll containing entity model, which would then create a SessionFactory, finally caching inside the computer memory. As the number of SessionFactory can be very large and will full of the memory, it has also devised a way of caching SessionFactory in the file. This post I hope this will not completely explain about and building a model of multiple databases. I just tried to mount a number of posts from the community and apply some of my knowledge to build a management model Session for ConfORM.As well as Fluent NHibernate, ConfORM also supported on the interface mapping, see this to understand it. So the first thing we will build the Entity Model for it, and here is what I will use the model for this article. A simple model for managing news and polls, it will be too easy for a number of people, but I hope not to bring complexity to this post.I will then have some code to build super type for the Entity Model. public interface IEntity<TId>    {        TId Id { get; set; }    } public abstract class EntityBase<TId> : IEntity<TId>    {        public virtual TId Id { get; set; }         public override bool Equals(object obj)        {            return Equals(obj as EntityBase<TId>);        }         private static bool IsTransient(EntityBase<TId> obj)        {            return obj != null &&            Equals(obj.Id, default(TId));        }         private Type GetUnproxiedType()        {            return GetType();        }         public virtual bool Equals(EntityBase<TId> other)        {            if (other == null)                return false;            if (ReferenceEquals(this, other))                return true;            if (!IsTransient(this) &&            !IsTransient(other) &&            Equals(Id, other.Id))            {                var otherType = other.GetUnproxiedType();                var thisType = GetUnproxiedType();                return thisType.IsAssignableFrom(otherType) ||                otherType.IsAssignableFrom(thisType);            }            return false;        }         public override int GetHashCode()        {            if (Equals(Id, default(TId)))                return base.GetHashCode();            return Id.GetHashCode();        }    } Database schema will be created as:The next step is to build the ConORM builder to create a NHibernate Configuration. Patrick have a excellent article about it at here. Contract of it below: public interface IConfigBuilder    {        Configuration BuildConfiguration(string connectionString, string sessionFactoryName);    } The idea here is that I will pass in a connection string and a set of the DLL containing the Entity Model and it makes me a NHibernate Configuration (shame that I stole this ideas of Sharp Architecture). And here is its code: public abstract class ConfORMConfigBuilder : RootObject, IConfigBuilder    {        private static IConfigurator _configurator;         protected IEnumerable<Type> DomainTypes;         private readonly IEnumerable<string> _assemblies;         protected ConfORMConfigBuilder(IEnumerable<string> assemblies)            : this(new Configurator(), assemblies)        {            _assemblies = assemblies;        }         protected ConfORMConfigBuilder(IConfigurator configurator, IEnumerable<string> assemblies)        {            _configurator = configurator;            _assemblies = assemblies;        }         public abstract void GetDatabaseIntegration(IDbIntegrationConfigurationProperties dBIntegration, string connectionString);         protected abstract HbmMapping GetMapping();         public Configuration BuildConfiguration(string connectionString, string sessionFactoryName)        {            Contract.Requires(!string.IsNullOrEmpty(connectionString), "ConnectionString is null or empty");            Contract.Requires(!string.IsNullOrEmpty(sessionFactoryName), "SessionFactory name is null or empty");            Contract.Requires(_configurator != null, "Configurator is null");             return CatchExceptionHelper.TryCatchFunction(                () =>                {                    DomainTypes = GetTypeOfEntities(_assemblies);                     if (DomainTypes == null)                        throw new Exception("Type of domains is null");                     var configure = new Configuration();                    configure.SessionFactoryName(sessionFactoryName);                     configure.Proxy(p => p.ProxyFactoryFactory<ProxyFactoryFactory>());                    configure.DataBaseIntegration(db => GetDatabaseIntegration(db, connectionString));                     if (_configurator.GetAppSettingString("IsCreateNewDatabase").ConvertToBoolean())                    {                        configure.SetProperty("hbm2ddl.auto", "create-drop");                    }                     configure.Properties.Add("default_schema", _configurator.GetAppSettingString("DefaultSchema"));                    configure.AddDeserializedMapping(GetMapping(),                                                     _configurator.GetAppSettingString("DocumentFileName"));                     SchemaMetadataUpdater.QuoteTableAndColumns(configure);                     return configure;                }, Logger);        }         protected IEnumerable<Type> GetTypeOfEntities(IEnumerable<string> assemblies)        {            var type = typeof(EntityBase<Guid>);            var domainTypes = new List<Type>();             foreach (var assembly in assemblies)            {                var realAssembly = Assembly.LoadFrom(assembly);                 if (realAssembly == null)                    throw new NullReferenceException();                 domainTypes.AddRange(realAssembly.GetTypes().Where(                    t =>                    {                        if (t.BaseType != null)                            return string.Compare(t.BaseType.FullName,                                          type.FullName) == 0;                        return false;                    }));            }             return domainTypes;        }    } I do not want to dependency on any RDBMS, so I made a builder as an abstract class, and so I will create a concrete instance for SQL Server 2008 as follows: public class SqlServerConfORMConfigBuilder : ConfORMConfigBuilder    {        public SqlServerConfORMConfigBuilder(IEnumerable<string> assemblies)            : base(assemblies)        {        }         public override void GetDatabaseIntegration(IDbIntegrationConfigurationProperties dBIntegration, string connectionString)        {            dBIntegration.Dialect<MsSql2008Dialect>();            dBIntegration.Driver<SqlClientDriver>();            dBIntegration.KeywordsAutoImport = Hbm2DDLKeyWords.AutoQuote;            dBIntegration.IsolationLevel = IsolationLevel.ReadCommitted;            dBIntegration.ConnectionString = connectionString;            dBIntegration.LogSqlInConsole = true;            dBIntegration.Timeout = 10;            dBIntegration.LogFormatedSql = true;            dBIntegration.HqlToSqlSubstitutions = "true 1, false 0, yes 'Y', no 'N'";        }         protected override HbmMapping GetMapping()        {            var orm = new ObjectRelationalMapper();             orm.Patterns.PoidStrategies.Add(new GuidPoidPattern());             var patternsAppliers = new CoolPatternsAppliersHolder(orm);            //patternsAppliers.Merge(new DatePropertyByNameApplier()).Merge(new MsSQL2008DateTimeApplier());            patternsAppliers.Merge(new ManyToOneColumnNamingApplier());            patternsAppliers.Merge(new OneToManyKeyColumnNamingApplier(orm));             var mapper = new Mapper(orm, patternsAppliers);             var entities = new List<Type>();             DomainDefinition(orm);            Customize(mapper);             entities.AddRange(DomainTypes);             return mapper.CompileMappingFor(entities);        }         private void DomainDefinition(IObjectRelationalMapper orm)        {            orm.TablePerClassHierarchy(new[] { typeof(EntityBase<Guid>) });            orm.TablePerClass(DomainTypes);             orm.OneToOne<News, Poll>();            orm.ManyToOne<Category, News>();             orm.Cascade<Category, News>(Cascade.All);            orm.Cascade<News, Poll>(Cascade.All);            orm.Cascade<User, Poll>(Cascade.All);        }         private static void Customize(Mapper mapper)        {            CustomizeRelations(mapper);            CustomizeTables(mapper);            CustomizeColumns(mapper);        }         private static void CustomizeRelations(Mapper mapper)        {        }         private static void CustomizeTables(Mapper mapper)        {        }         private static void CustomizeColumns(Mapper mapper)        {            mapper.Class<Category>(                cm =>                {                    cm.Property(x => x.Name, m => m.NotNullable(true));                    cm.Property(x => x.CreatedDate, m => m.NotNullable(true));                });             mapper.Class<News>(                cm =>                {                    cm.Property(x => x.Title, m => m.NotNullable(true));                    cm.Property(x => x.ShortDescription, m => m.NotNullable(true));                    cm.Property(x => x.Content, m => m.NotNullable(true));                });             mapper.Class<Poll>(                cm =>                {                    cm.Property(x => x.Value, m => m.NotNullable(true));                    cm.Property(x => x.VoteDate, m => m.NotNullable(true));                    cm.Property(x => x.WhoVote, m => m.NotNullable(true));                });             mapper.Class<User>(                cm =>                {                    cm.Property(x => x.UserName, m => m.NotNullable(true));                    cm.Property(x => x.Password, m => m.NotNullable(true));                });        }    } As you can see that we can do so many things in this class, such as custom entity relationships, custom binding on the columns, custom table name, ... Here I only made two so-Appliers for OneToMany and ManyToOne relationships, you can refer to it here public class ManyToOneColumnNamingApplier : IPatternApplier<PropertyPath, IManyToOneMapper>    {        #region IPatternApplier<PropertyPath,IManyToOneMapper> Members         public void Apply(PropertyPath subject, IManyToOneMapper applyTo)        {            applyTo.Column(subject.ToColumnName() + "Id");        }         #endregion         #region IPattern<PropertyPath> Members         public bool Match(PropertyPath subject)        {            return subject != null;        }         #endregion    } public class OneToManyKeyColumnNamingApplier : OneToManyPattern, IPatternApplier<PropertyPath, ICollectionPropertiesMapper>    {        public OneToManyKeyColumnNamingApplier(IDomainInspector domainInspector) : base(domainInspector) { }         #region Implementation of IPattern<PropertyPath>         public bool Match(PropertyPath subject)        {            return Match(subject.LocalMember);        }         #endregion Implementation of IPattern<PropertyPath>         #region Implementation of IPatternApplier<PropertyPath,ICollectionPropertiesMapper>         public void Apply(PropertyPath subject, ICollectionPropertiesMapper applyTo)        {            applyTo.Key(km => km.Column(GetKeyColumnName(subject)));        }         #endregion Implementation of IPatternApplier<PropertyPath,ICollectionPropertiesMapper>         protected virtual string GetKeyColumnName(PropertyPath subject)        {            Type propertyType = subject.LocalMember.GetPropertyOrFieldType();            Type childType = propertyType.DetermineCollectionElementType();            var entity = subject.GetContainerEntity(DomainInspector);            var parentPropertyInChild = childType.GetFirstPropertyOfType(entity);            var baseName = parentPropertyInChild == null ? subject.PreviousPath == null ? entity.Name : entity.Name + subject.PreviousPath : parentPropertyInChild.Name;            return GetKeyColumnName(baseName);        }         protected virtual string GetKeyColumnName(string baseName)        {            return string.Format("{0}Id", baseName);        }    } Everyone also can download the ConfORM source at google code and see example inside it. Next part I will write about multiple database factory. Hope you enjoy about it. happy coding and see you next part.

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  • Microsoft 2003 DNS sometimes cant query for some A pointers when their TTL expires

    - by Bq
    Warning Long question :) We have a win 2003 server with a DNS server, every now and then it cant provide us with some A pointers for a specific domain. I have a small script running which asks for SOA,NS and A records for the domain in question and sometimes when the TTL expires the DNS fails to get the A records again, a Clear Cache fixes the problem.. Have a look Here it worked when the TTL expired Thu Apr 29 15:24:20 METDST 2010 dig basefarm.net soa basefarm.net. 64908 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 64908 IN NS ns01.sth.basefarm.net. basefarm.net. 64908 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 299 IN A 80.76.149.76 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 299 IN A 80.76.149.76 The TTL expired for ns01.sth.basefarm.net and ns01.osl.basefarm.net but the DNS managed to get the new values (TTL 3600) Thu Apr 29 15:29:20 METDST 2010 dig basefarm.net soa basefarm.net. 64608 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 64608 IN NS ns01.sth.basefarm.net. basefarm.net. 64608 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 3600 IN A 80.76.149.76 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 3600 IN A 80.76.149.76 But then another time it fails, and we need to clear the dns cache for it to start working again... Thu Apr 29 17:24:23 METDST 2010 dig basefarm.net soa basefarm.net. 57705 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 57705 IN NS ns01.sth.basefarm.net. basefarm.net. 57705 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 299 IN A 80.76.149.76 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 299 IN A 80.76.149.76 The TTL expires but the DNS cant get the ip addresses for ns01.sth.basefarm.net and ns01.osl.basefarm.net Thu Apr 29 17:29:23 METDST 2010 dig basefarm.net soa basefarm.net. 57405 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 57405 IN NS ns01.sth.basefarm.net. basefarm.net. 57405 IN NS ns01.osl.basefarm.net. dig ns01.sth.basefarm.net a Lookup failed I'm really lost on this one and have tried asking Google but to no avail..

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  • Can't get MySQL to install

    - by James Marthenal
    I'd like to think I know what I'm doing in a Unix shell but maybe not. I made a mistake in a configuration file for MySQL, so I decided to just uninstall it and then reinstall it, so I did: sudo apt-get --purge remove mysql-server mysql-server-5.0 mysql-client The files were deleted, so I then tried to install it, but it didn't ask me for a root password or anything else, so I uninstalled it using the above command again and then did sudo rm -rf /etc/mysql sudo rm /etc/init.d/mysql sudo rm -rf /var/lib/mysql* I then restarted the computer then installed it again: sudo apt-get install mysql-server mysql-client It asked for a root password, and everything looked like it would work, until I saw this: $ sudo apt-get install mysql-server mysql-client Reading package lists... Done Building dependency tree Reading state information... Done The following extra packages will be installed: mysql-server-5.0 Suggested packages: tinyca The following NEW packages will be installed: mysql-client mysql-server mysql-server-5.0 0 upgraded, 3 newly installed, 0 to remove and 1 not upgraded. Need to get 0B/27.4MB of archives. After this operation, 86.7MB of additional disk space will be used. Do you want to continue [Y/n]? y WARNING: The following packages cannot be authenticated! mysql-server-5.0 mysql-client mysql-server Authentication warning overridden. Preconfiguring packages ... Can't exec "/tmp/mysql-server-5.0.config.28101": Permission denied at /usr/share/perl/5.10/IPC/Open3.pm line 168. open2: exec of /tmp/mysql-server-5.0.config.28101 configure failed at /usr/share/perl5/Debconf/ConfModule.pm line 59 mysql-server-5.0 failed to preconfigure, with exit status 255 Selecting previously deselected package mysql-server-5.0. (Reading database ... 160284 files and directories currently installed.) Unpacking mysql-server-5.0 (from .../mysql-server-5.0_5.0.51a-24+lenny5_amd64.deb) ... Selecting previously deselected package mysql-client. Unpacking mysql-client (from .../mysql-client_5.0.51a-24+lenny5_all.deb) ... Selecting previously deselected package mysql-server. Unpacking mysql-server (from .../mysql-server_5.0.51a-24+lenny5_all.deb) ... Processing triggers for man-db ... Setting up mysql-server-5.0 (5.0.51a-24+lenny5) ... Stopping MySQL database server: mysqld. /var/lib/dpkg/info/mysql-server-5.0.postinst: line 144: /etc/mysql/conf.d/old_passwords.cnf: No such file or directory dpkg: error processing mysql-server-5.0 (--configure): subprocess post-installation script returned error exit status 1 Setting up mysql-client (5.0.51a-24+lenny5) ... dpkg: dependency problems prevent configuration of mysql-server: mysql-server depends on mysql-server-5.0; however: Package mysql-server-5.0 is not configured yet. dpkg: error processing mysql-server (--configure): dependency problems - leaving unconfigured Errors were encountered while processing: mysql-server-5.0 mysql-server E: Sub-process /usr/bin/dpkg returned an error code (1) Now I can't seem to figure out what to do. I just want to get a clean MySQL installation at this point. I'm running the latest stable release of Debian. All help is appreciated—thanks! Edit: I looked at this similar question, which suggests that I uninstall mysql-common, but when I try to do so I see: The following packages will be REMOVED: apache2 apache2-mpm-prefork apache2-utils apache2.2-common git-svn libapache2-mod-php5 libapache2-mod-python libapache2-svn libaprutil1 libdbd-mysql-perl libdbd-mysql-rubygem libmysql-ruby libmysql-ruby1.8 libmysql-rubygem libmysqlclient15-dev libmysqlclient15off librdf-perl librdf0 libserf-0-0 libsvn-perl libsvn1 mysql-client-5.0 mysql-common mytop ndn-apache22-php5 ndn-apache22-svn ndn-interpreters ndn-lighttpd ndn-netsaint-plugins ndn-perl-modules ndn-php5-cgi ndn-php5-xcache ndn-php53 ndn-php53-suhosin ndn-rubygems php5 php5-mcrypt php5-mysql proftpd proftpd-mod-mysql python-django python-mysqldb python-subversion python-svn subversion subversion-tools trac zendoptimizer 0 upgraded, 0 newly installed, 48 to remove and 1 not upgraded. Eeek! Any suggestions?

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  • Rails requires Rubygems but I have the gems

    - by fogonthedowns
    Update I notice that which ruby and whereis ruby are different locations which ruby /opt/local/bin/ruby whereis ruby /usr/bin/ruby I recently upgraded ruby to ruby 1.8.7 (2009-06-12 patchlevel 174) [i686-darwin10] and I think I broke rails. When I attempt to load rails. I get an odd message. Please help! $ ruby script/server Rails requires RubyGems = 1.3.2. Please install RubyGems and try again: http://rubygems.rubyforge.org $ rails -v Rails 3.0.0.beta $ gem -v 1.3.6 $ which gem /usr/bin/gem $ whereis gem /usr/bin/gem $ which rails /usr/bin/rails $ whereis rails /usr/bin/rails $ /usr/bin/gem -v 1.3.6 $ /usr/bin/rails -v Rails 3.0.0.beta $ ruby script/console Rails requires RubyGems >= 1.3.2. Please install RubyGems and try again: http://rubygems.rubyforge.org $ gem list rails *** LOCAL GEMS *** rails (3.0.0.beta, 2.3.5, 2.2.2, 1.2.6) $ gem list *** LOCAL GEMS *** abstract (1.0.0) actionmailer (3.0.0.beta, 2.3.5, 2.2.2, 1.3.6) actionpack (3.0.0.beta, 2.3.5, 2.2.2, 1.13.6) actionwebservice (1.2.6) activemerchant (1.4.1) activemodel (3.0.0.beta) activerecord (3.0.0.beta, 2.3.5, 2.2.2, 1.15.6) activerecord-tableless (0.1.0) activeresource (3.0.0.beta, 2.3.5, 2.2.2) activesupport (3.0.0.beta, 2.3.5, 2.2.2, 1.4.4) acts_as_ferret (0.4.3) arel (0.2.pre) authlogic (2.1.3) builder (2.1.2) bundler (0.9.3) calendar_date_select (1.15) capistrano (2.5.2) cgi_multipart_eof_fix (2.5.0) chronic (0.2.3) columnize (0.3.1) compass (0.8.17) daemons (1.0.10) dnssd (0.6.0) erubis (2.6.5) fastercsv (1.5.0) fastthread (1.0.1) fcgi (0.8.7) ferret (0.11.6) flay (1.4.0) flog (2.4.0) gbarcode (0.98.16) gem_plugin (0.2.3) git (1.2.5) haml (2.2.15) haml-edge (2.3.100) highline (1.5.0) hoe (2.4.0) hpricot (0.6.164) i18n (0.3.3) javan-whenever (0.3.7) jeweler (1.4.0) jscruggs-metric_fu (1.1.5) json_pure (1.2.0) libxml-ruby (1.1.2) linecache (0.43) mail (2.1.2) mechanize (0.9.3) memcache-client (1.7.8) mime-types (1.16) mislav-will_paginate (2.3.11) mocha (0.9.7) mojombo-chronic (0.3.0) mongrel (1.1.5) needle (1.3.0) net-scp (1.0.1) net-sftp (2.0.1, 1.1.1) net-ssh (2.0.4, 1.1.4) net-ssh-gateway (1.0.0) nifty-generators (0.3.0) nokogiri (1.4.0) openrain-action_mailer_tls (1.1.3) passenger (2.2.5) polyglot (0.2.9) prawn (0.6.3) prawn-core (0.6.3) prawn-format (0.2.3) prawn-layout (0.3.2) prawn-security (0.1.1) rack (1.1.0, 1.0.1) rack-mount (0.4.5) rack-test (0.5.3) rails (3.0.0.beta, 2.3.5, 2.2.2, 1.2.6) railties (3.0.0.beta) rake (0.8.7, 0.8.3) rake-compiler (0.6.0) RedCloth (4.1.1) reek (1.2.6) relevance-rcov (0.9.2.1) rmagick (2.12.2) roodi (2.1.0) rsl-stringex (1.0.3) rspec (1.2.9) rspec-rails (1.2.9) ruby-debug (0.10.3) ruby-debug-base (0.10.3) ruby-openid (2.1.2) ruby-yadis (0.3.4) ruby2ruby (1.2.4) ruby_parser (2.0.4) rubyforge (2.0.3) rubygems-update (1.3.6, 1.3.5) rubynode (0.1.5) searchlogic (2.3.9) sexp_processor (3.0.3) spree (0.9.4) sqlite3-ruby (1.2.5, 1.2.4) termios (0.9.4) test-unit (2.0.5) text-format (1.0.0) text-hyphen (1.0.0) thor (0.13.0) tlsmail (0.0.1) topfunky-gruff (0.3.5) treetop (1.4.3) tzinfo (0.3.16) xmpp4r (0.4)

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  • Debian, 6rd tunnel, and connection troubles

    - by Chris B
    Long story short I am having issues with IPv6 using a 6rd tunnel with my ISP, charter business. They offer a 6rd tunnel that I think I have properly set up, but the server doesn’t reply to every ipv6 request. When the server has the network interfaces idle with no traffic for about 10 minutes, then IPv6 stops accepting inbound connections. to re-allow it, I must go into the server, and make it do a outbound ipv6 connection (normally a ping) to start it back up. Whats weird though i that if I run iptraf when its not working, it still shows a inbound ipv6 packet… the server is just not replying, and I can’t figure out why. Also, if I try to access my server over IPv6 from a house about 1 mile away on the same ISP, it is never able to connect. it always times out, but again the iptraf shows a ipv6 inbound packet. Again, it just does not reply. To test if my server is accessible through IPv6 I always have to use my vzw 4g phone (they use IPv6) or ipv6proxy dot net. Here is all of the configuration information my ISP gives on there tunnel server: 6rd Prefix = 2602:100::/32 Border Relay Address = 68.114.165.1 6rd prefix length = 32 IPv4 mask length = 0 Here is my /etc/network/interfaces for ipv6 (used x's to block real addresses) auto charterv6 iface charterv6 inet6 v4tunnel address 2602:100:189f:xxxx::1 netmask 32 ttl 64 gateway ::68.114.165.1 endpoint 68.114.165.1 local 24.159.218.xxx up ip link set mtu 1280 dev charterv6 here is my iptables config filter :INPUT DROP [0:0] :fail2ban-ssh – [0:0] :OUTPUT ACCEPT [0:0] :FORWARD DROP [0:0] :hold – [0:0] -A INPUT -p tcp -m tcp —dport 22 -j fail2ban-ssh -A INPUT -m state —state RELATED,ESTABLISHED -j ACCEPT -A INPUT -p tcp -m multiport -j ACCEPT —dports 80,443,25,465,110,995,143,993,587,465,22 -A INPUT -i lo -j ACCEPT -A INPUT -p tcp -m tcp —dport 10000 -j ACCEPT -A INPUT -p tcp -m tcp —dport 5900:5910 -j ACCEPT -A fail2ban-ssh -j RETURN -A INPUT -p icmp -j ACCEPT COMMIT and last here is my ip6tables firewall config filter :INPUT DROP [1653:339023] :FORWARD DROP [0:0] :OUTPUT ACCEPT [60141:13757903] :hold – [0:0] -A INPUT -m state —state RELATED,ESTABLISHED -j ACCEPT -A INPUT -p tcp -m multiport —dports 80,443,25,465,110,995,143,993,587,465,22 -j ACCEPT -A INPUT -i lo -j ACCEPT -A INPUT -p tcp -m tcp —dport 10000 -j ACCEPT -A INPUT -p tcp -m tcp —dport 5900:5910 -j ACCEPT -A INPUT -p ipv6-icmp -j ACCEPT COMMIT So Summary: 1.iptraf always shows IPv6 traffic, so its always making it to the server 2.server stops replying on ipv6 after no traffic for awhile (10 minutesish) until a outbound connection is made, then the process repeats. 3.server is NEVER accessable vi same ISP (yet iptraf still shows ipv6 request) Notes: When I try to access it from the same ISP from across town, even with iptables and ip6tables allowing ALL inbound traffic, this is what iptraf shows. IPv6 (92 bytes) from 97.92.18.xxx to 24.159.218.xxx on eth0 ICMP dest unrch (port) (120 bytes) from 24.159.218.xxx to 97.92.18.xxx on eth1 its strange, like its trying to forward to LAN? (eth1 is LAN, eth0 is WAN) even with the IPv6 address being set in the hosts file to the servers domain name. With iptables set up normally with the above configurations it only says this: IPv6 (100 bytes) from 97.92.18.xxx to 24.159.218.xxx on eth0 Im REALLY stuck on this, and any help would be GREATLY appreciated.

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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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  • DCOM Authentication Fails to use Kerberos, Falls back to NTLM

    - by Asa Yeamans
    I have a webservice that is written in Classic ASP. In this web service it attempts to create a VirtualServer.Application object on another server via DCOM. This fails with Permission Denied. However I have another component instantiated in this same webservice on the same remote server, that is created without problems. This component is a custom-in house component. The webservice is called from a standalone EXE program that calls it via WinHTTP. It has been verified that WinHTTP is authenticating with Kerberos to the webservice successfully. The user authenticated to the webservice is the Administrator user. The EXE to webservice authentication step is successful and with kerberos. I have verified the DCOM permissions on the remote computer with DCOMCNFG. The default limits allow administrators both local and remote activation, both local and remote access, and both local and remote launch. The default component permissions allow the same. This has been verified. The individual component permissions for the working component are set to defaults. The individual component permissions for the VirtualServer.Application component are also set to defaults. Based upon these settings, the webservice should be able to instantiate and access the components on the remote computer. Setting up a Wireshark trace while running both tests, one with the working component and one with the VirtualServer.Application component reveals an intresting behavior. When the webservice is instantiating the working, custom, component, I can see the request on the wire to the RPCSS endpoint mapper first perform the TCP connect sequence. Then I see it perform the bind request with the appropriate security package, in this case kerberos. After it obtains the endpoint for the working DCOM component, it connects to the DCOM endpoint authenticating again via Kerberos, and it successfully is able to instantiate and communicate. On the failing VirtualServer.Application component, I again see the bind request with kerberos go to the RPCC endpoing mapper successfully. However, when it then attempts to connect to the endpoint in the Virtual Server process, it fails to connect because it only attempts to authenticate with NTLM, which ultimately fails, because the webservice does not have access to the credentials to perform the NTLM hash. Why is it attempting to authenticate via NTLM? Additional Information: Both components run on the same server via DCOM Both components run as Local System on the server Both components are Win32 Service components Both components have the exact same launch/access/activation DCOM permissions Both Win32 Services are set to run as Local System The permission denied is not a permissions issue as far as I can tell, it is an authentication issue. Permission is denied because NTLM authentication is used with a NULL username instead of Kerberos Delegation Constrained delegation is setup on the server hosting the webservice. The server hosting the webservice is allowed to delegate to rpcss/dcom-server-name The server hosting the webservice is allowed to delegate to vssvc/dcom-server-name The dcom server is allowed to delegate to rpcss/webservice-server The SPN's registered on the dcom server include rpcss/dcom-server-name and vssvc/dcom-server-name as well as the HOST/dcom-server-name related SPNs The SPN's registered on the webservice-server include rpcss/webservice-server and the HOST/webservice-server related SPNs Anybody have any Ideas why the attempt to create a VirtualServer.Application object on a remote server is falling back to NTLM authentication causing it to fail and get permission denied? Additional information: When the following code is run in the context of the webservice, directly via a testing-only, just-developed COM component, it fails on the specified line with Access Denied. COSERVERINFO csi; csi.dwReserved1=0; csi.pwszName=L"terahnee.rivin.net"; csi.pAuthInfo=NULL; csi.dwReserved2=NULL; hr=CoGetClassObject(CLSID_VirtualServer, CLSCTX_ALL, &csi, IID_IClassFactory, (void **) &pClsFact); if(FAILED( hr )) goto error1; // Fails here with HRESULT_FROM_WIN32(ERROR_ACCESS_DENIED) hr=pClsFact->CreateInstance(NULL, IID_IUnknown, (void **) &pUnk); if(FAILED( hr )) goto error2; Ive also noticed that in the Wireshark Traces, i see the attempt to connect to the service process component only requests NTLMSSP authentication, it doesnt even attmept to use kerberos. This suggests that for some reason the webservice thinks it cant use kerberos...

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  • Lag spikes at full CPU usage, lagy mouse, maybe video card

    - by Roberts
    My PC specs: Motherboard Name - Gigabyte GA-945PL-S3 CPU Type - DualCore Intel Core 2 Duo E4300, 1800 MHz (9 x 200) OS - Microsoft Windows 7 Ultimate OS Kernel Type - 32-bit OS Version - 6.1.7601 I bougth a new video card one month ago. GeForce 210. I didn't have any problems. I wanted to overclock it, in other words: "Play with it". So I installed Gigabyte EasyBoost from CD and overclocked the GPU 590 + 110 mhz, memory to max to 960mhz from 800mhz. Benchmarks showed a little bit bigger score. Then I overclocked shader clock from 1405 to [..] (don't remeber really). So I was playing Modern Warfare 2 when off sudden computer froze when I wanted to select team, I was afk before that. I had to reset CMOS. After that I had problems with Skype: unread messages and no sound. Then I figured it out that when ever I open EasyBoost - Skype starts to glitch again. Now I use EVGA Precission X. Now after a month, I cleaned computer and closed the case, it was open all the time. I started to overclock GPU clock only (just a bit) because there was no problems that would stop me. So sometimes on heavy CPU load graphics starts to lag. Dragging a window is painful to watch too. Sometimes the screen freezes for 5 to 10 seconds (I can see that hard disk activity is maximal). You may say that CPU fault it is, isn't it? But sometimes lag spikes starts randomly when CPU load is at maximum. All 3 benchmark softwares (PerformanceTest, NovaBench and MSI Kombustor) shows that performance of my video card has dropped about 25%. BUT! CPU score is lower too. I ignored these problems but when I refreshed Windows Experience Index I was shocked. Month before (in latvian language but not so hard to understand): Now 01.04.2012 (upgraded RAM): This happened when I tried to capture Minecraft with Fraps on underclocked GPU to 580mhz (def: 590mhz): All drivers are up to date. Average CPU temperature from 55°C to 75°C (at 70°C sometimes starts these lag spikes). Video card's tempratures are from 45°C to 60°C (very hard to reach 60°C). So my hope is that the video card is fine, cause this card is very new and I want to upgrade CPU anyways. Aplogies for my mistakes in vocabulary (I am trying to type this as fast I can). Update 02.04.2012 - 7:21 Forgot one thing, my hard disk is extrimly slow and I will upgrade it this week or next week so I will be installing same OS again. I am multi-tasker but I can't do much because of 1.8 GHz CPU and slow hard drive (Model ID - WDC WD800JD-60JRC0). The Windows Experience Index is back to normal. Actually "Spelu grafika" (Gaming graphics) are higher than month ago. During this test mouse was very lagy, but month ago there weren't any problems. WHY!?

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  • Why do GPUs overheat?

    - by JAD
    About a year ago, I added a 9800GT (1 GB version) and a Corsair CX500 PSU to an HP M8000N computer. A few weeks ago, the HDD overheated and I decided to transfer the GPU & PSU to a new build, which consists of: i3 @ 3.3Ghz Gigabyte H61 Micro ATX Mobo 4GB RAM 500GB WD HDD DVD RW Drive Cooler Master Elite 430 Tower Once I had Win7 up and running, I installed all the essential drivers that came with the Gigabyte Mobo CD. However, whenever I tried installing the Graphics Media Accelerator driver, the computer would crash and enter an endless boot sequence on the next startup. I skipped installing this driver and installed the CD driver for the 9800GT, which by now is a year old. Everything was working fine, WEI rated my GPU at 6.6 graphics & aero performance. However, after updating my Nvidia drivers to the latest, the WEI dropped my rating to 3.3 for Aero, and 4.7 for graphics performance. Just to make sure that everything was ok, I ran Bad Company 2 on medium settings. The first few minutes ran just fine at a smooth framerate, so I dismissed this as Windows being Windows. About 6 hours later, I ran BC2 again. This time I averaged anywhere from 2-5 FPS. I checked the GPU temperature through GPU-Z, and it came back as 120C. The problem with this, is that the computer was on for six hours up to that point. Wouldn't the card have experienced a reactor core meltdown a lot sooner than that? Granted, the computer was "sleeping" some of the time, but still... The next day I took out a temperature gun and ran some tests. I would point the laser at a very specific area on the reverse side of the card (not the fan or "front"), and compare the temp reading with GPU-Z. After leaving the system on idle on idle for a few minutes, I ran BC2 twice. Here are the results: GPU-Z Reading / Temp Gun Reading / Time Null / 22.3°C / Comp is Off 53°C / 33.5°C / 1:49 78°C / 46°C / 1:53 - (First BC2 run; good framerate) 102°C / 64.6°C / 2:01 - (System is again on idle) 113°C / 64.8°C / 2:10 119°C / 71.8°C / 2:17 - (Second BC2 run; poor framerate) I should also mention that I also took a temp recording of another part of the GPU from 2:01-2:17. The temp in this area jumped from 75°C to 82.9°C in that time frame. This pretty much confirms that GPU-Z is reporting the temperature accurately, and the card is overheating. But I'd like to know why; the cars is doing nothing and still the temperature climbs at a steady rate. I thoroughly cleaned the GPU and PSU when I salvaged them from the old HP M8000N computer with a can of compressed air, dust cant be the issue. Similarly, the rest of the computer is brand new. I installed various Nvidia drivers, but no luck. It seems strange to me that a year-old card is suddenly failing on me; aren't they supposed to last at least two years? Could this be a driver issue? Is the motherboard faulty? Could the PSU be overfeeding the card on voltage? Neither case seems likely, as the CPU, RAM and otherwise the rest of the comp has worked flawlessly and has stayed well within respectable temp ranges (the i3 lingers around 50C, the HDD stays at 30C, so does the PSU). How can I pinpoint the issue?

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  • Netgear VPN endpoint drops connectivity to single IP address

    - by Justin Bowers
    I'm having a strange issue with one of the networks I manage recently. We have about 14 different networks connected together through a Netgear hardware VPN. Everything has been running fine (other than standard connectivity problems) for a few years now, but I've hit a wall with a problem that's just cropped up at one of the VPN endpoint locations. Our primary VPN network is on the 192.168.1.0/24 subnet and our other 13 networks are on the 192.168.2.0/24 - 192.168.14.0/24 subnets. We run a terminal server on the 192.168.1.0/24 network with IP address 192.168.1.100. Starting Thursday of last week, we had a problem with connectivity of the 192.168.2.0/24 network to 192.168.1.100. When troubleshooting the problem, I found that Network 2 (192.168.2.0/24) still had connectivity to the Internet as well as VPN connectivity to Network 1 (192.168.1.0/24). We could ping and connect to any other device other than the server with IP address 192.168.1.100. Also, none of our networks had an issue accessing 192.168.1.100. I ran a scan on Network 2 after assigning static IP addresses to one of the workstations but received no response from 192.168.1.100 (looking for possibly a new device that someone had plugged into Network 2 that had a duplicate IP address with the server). Asking the staff, noone had reported connecting a new device to Network 2 as well. I then assigned a secondary IP address of 192.168.1.88 to the server and could ping and connect to the secondary IP address from Network 2, but still couldn't access it via 192.168.1.100. I then just rebooted the Netgear VPN Firewall (FVS318v3) and after it came back up, connectivity to 192.168.1.100 was restored. Beforehand, when checking for devices with a possible duplicate IP address, I did run a check for available wireless access points and stations and found none (our wireless is secured via MAC address access control through a WG102 device). I thought that it may have been a fluke for some reason since everything came back up after a power cycle of the VPN Firewall. Things ran fine for a few days until this afternoon, when the problem happened again. One of our users claimed that they had connectivity problems to the server and after connecting to the computer, I found that I couldn't ping the server address anymore. I could still ping the alternate IP address of the server though, so I went ahead and rebooted the VPN firewall again and connectivity was restored. Unfortunately, I can't find anything in the security or VPN logs of the firewall that helps point me in the right direction, so I thought I would go ahead and ask to see if anyone else has any other insight into why we've started having this problem. I am aware that it could still be a device with a duplicate IP address of the server on Network 2, but every employee claim states that there's been no such new device brought in to the network. I know this is a long read, but any help is appreciated! Thanks, Justin

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  • Why wireless adatper stop to work?

    - by AndreaNobili
    today I correctly installed the driver for the TP-LINK TL-WN725N USB wireless adapter on my RaspBerry Pi (I use RaspBian that is a Debian), then I setted up the wifi using the wpa-supplicant as explained in this tutorial: http://www.maketecheasier.com/setup-wifi-on-raspberry-pi/ This worked fine untill this evening. Then suddenly it stopped to work when I try to connect in SSH and the Raspberry is on the wireless (or rather it should be, as this is not in the list of my router's DHCP connected Client) The strange thing is that the USB wirless adapter blink so I think that this is not a driver problem. If I try to connect it by the ethernet I have no problem. It appear in my router's DHCP connected Client and I can connect to it by SSH. When I connect to it using ethernet if I perform an ifconfig command I obtain: pi@raspberrypi ~ $ ifconfig eth0 Link encap:Ethernet HWaddr b8:27:eb:2a:9f:b0 inet addr:192.168.1.9 Bcast:192.168.1.255 Mask:255.255.255.0 UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:48 errors:0 dropped:0 overruns:0 frame:0 TX packets:59 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:6006 (5.8 KiB) TX bytes:8268 (8.0 KiB) lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 UP LOOPBACK RUNNING MTU:65536 Metric:1 RX packets:8 errors:0 dropped:0 overruns:0 frame:0 TX packets:8 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:1104 (1.0 KiB) TX bytes:1104 (1.0 KiB) wlan0 Link encap:Ethernet HWaddr e8:94:f6:19:80:4c UP BROADCAST MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) So it seems that the wlan0 USB wireless adapter driver is correctly loaded. If I remove the USB wireless adapter and put it again into the USB port, the lasts lines of dmesg log is: [ 20.303172] smsc95xx 1-1.1:1.0 eth0: hardware isn't capable of remote wakeup [ 20.306340] RTL871X: set bssid:00:00:00:00:00:00 [ 20.306726] RTL871X: set ssid [g\xffffffc6isQ\xffffffffJ\xffffffec)\xffffffcd\xffffffba\xffffffba\xffffffab\xfffffff2\xfffffffb\xffffffe3F|\xffffffc2T\xfffffff8\x1b\xffffffe8\xffffffe7\xffffff8dvZ.c3\xffffff9f\xffffffc9\xffffff9a\xffffff9aD\xffffffa7\x1a\xffffffa0\x1a\xffffff8b] fw_state=0x00000008 [ 21.614585] RTL871X: indicate disassoc [ 21.908495] smsc95xx 1-1.1:1.0 eth0: link up, 100Mbps, full-duplex, lpa 0x45E1 [ 25.006282] Adding 102396k swap on /var/swap. Priority:-1 extents:1 across:102396k SSFS [ 26.247997] RTL871X: nolinked power save enter As you can see some of these line are related to the RTL871X that is my USB wireless adapter, but I don't know is that these line report an error or if it is all ok. Looking at the adapter status I obtain: pi@raspberrypi ~ $ ip link list dev wlan0 3: wlan0: <NO-CARRIER,BROADCAST,MULTICAST,UP> mtu 1500 qdisc mq state DOWN mode DORMANT qlen 1000 link/ether e8:94:f6:19:80:4c brd ff:ff:ff:ff:ff:ff As you can see the mode is DORMANT but I think that this is normal because now I am connected using ethernet. I tryied to set up the adapter but it seems that I obtain no result, infact: pi@raspberrypi ~ $ sudo ip link set dev wlan0 up pi@raspberrypi ~ $ ip link list dev wlan0 3: wlan0: <NO-CARRIER,BROADCAST,MULTICAST,UP> mtu 1500 qdisc mq state DOWN mode DORMANT qlen 1000 link/ether e8:94:f6:19:80:4c brd ff:ff:ff:ff:ff:ff pi@raspberrypi ~ $ sudo ip link set dev wlan0 up This is my /etc/network/interfaces file content and it is ok: auto lo iface lo inet loopback iface eth0 inet dhcp allow-hotplug wlan0 iface wlan0 inet manual wpa-roam /etc/wpa_supplicant/wpa_supplicant.conf iface default inet dhcp and it is the /etc/wpa_supplicant/wpa_supplicant.conf that I think is ok (I did not change it compared to when it worked): ctrl_interface=DIR=/var/run/wpa_supplicant GROUP=netdev update_config=1 network={ ssid="MY-NETWORK" psk="mypassword" key_mgmt=WPA-PSK } and infact if I execute a network scan I correctly find MY-NETWORK in the network list,infact: pi@raspberrypi ~ $ sudo iwlist wlan0 scan | grep ESSID ESSID:"TeleTu_74888B0060AD" ESSID:"MY-NETWORK" ESSID:"FASTWEB-1-PT6NtjL4TOSe" ESSID:"DC" So I reboot the system and I remove the ethernet cable but when I try to connect again to my raspberry I obatin the following error message: andrea@andrea-virtual-machine:~$ sudo ssh [email protected] ssh: connect to host 192.168.1.9 port 22: No route to host It seems that it can't connect using wireless. What could be the problem? What am I missing? How can I solve this situation? Tnx

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  • MS Securily Essentials efficiency / usage, suspicious processes

    - by biggvsdiccvs
    I recently noticed that my (originally pretty fast) Windows 7 Pro laptop started getting slow and using a lot of CPU power for no apparent reason. A full scan by Microsoft Security Essentials revealed nothing. After some investigation, I found multiple instances of a strange process called urpev.exe and a couple of similar exe files sitting in subdirectories of Users//AppData/Roaming (this particular one was in a folder called Xyceowme). Description: "Mescrosift Visaal Studie 2010". Company name: "Mesrosift Corporatien". Is it a virus or something? :) Now, all of these exe files were scheduled to be started from the Task Scheduler by tasks with names like "Security Center Update - 1291373911" and similar. My user name was listed as the author of the tasks. I disabled the tasks, restarted the computer in safe mode and moved all of the exe files to quarantine for further investigation. All of this was done last night. I just scanned the files with Security Essentials again (not updated since yesterday) in the quarantine location and this time it found PWS:Win32/Zbot.gen!plock in urpev.exe (but not in the other exe files, which are most likely viruses, too). Category: Password Stealer Description: This program is dangerous and captures user passwords. Another strange process is browser.exe (not chrome.exe) by Google Inc., described as Google Chrome. I uninstalled Chrome but it's still there. It runs out of Users\\AppData\LocalLow\UIVoice\ToolMedium\browser.exe and if I move it in safe mode, it just reappears there, and multiple instances run. Needless to say, it I kill it, it just runs again. Couldn't see anything in Task Scheduler, but found a couple of references to it in the Registry Editor: HKEY_CURRENT_USER/Software/Microsoft/Internet Explorer/LowRegistry/Audio/PolicyConfig/PropertyStore/ HKEY_USERS/S-1-5-21-1685709306-872053864-2599010960-1002/Software/Microsoft/Internet Explorer/LowRegistry/Audio/PolicyConfig/PropertyStore/ Maybe it's a legit process, but seems kind of strange. For the time being, I suspended the process and killed all of the child processes when I booted up the laptop. I used Security Essentials to scan the system periodically, but obviously it's not effective at least against one virus. I had the "real-time protection" turned off. Would it help if it were turned on and how much of a nuisance would it be? I wonder if there is a better alternative to Security Essentials. Over the years I've used multiple antivirus products at home and especially at work and was not very happy with any of them. Apparently, asking for software recommendations or comparisons is taboo here, but I will mention that I installed Malware Bytes and it was able to find an quarantine a bunch of suspicious files, and at least some of which were truly infected, but when it scans the bogus security center update executables from Mesrosift Corporatien, it finds nothing wrong. Also, any thoughts on the browser.exe mystery? Neither MS Security Essentials nor Malware Bytes found anything wrong with that file. However, after I ran a Malware Bytes scan and quarantined everything it found suspicious and rebooted the laptop, the process did not run.

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  • Linux not buffering block I/O when the device is not "in use" (i.e. mounted)

    - by Radek Hladík
    I am installing new server and I've found an interesting issue. The server is running Fedora 19 (3.11.7-200.fc19.x86_64 kernel) and is supposed to host a few KVM/Qemu virtual servers (mail server, file server, etc..). The HW is Intel(R) Xeon(R) CPU 5160 @ 3.00GHz with 16GB RAM. One of the most important features will be Samba server and we have decided to make it as virtual machine with almost direct access to the disks. So the real HDD is cached on SSD (via bcache) then raided with md and the final device is exported into the virtual machine via virtio. The virtual machine is again Fedora 19 with the same kernel. One important topic to find out is whether the virtualization layer will not introduce high overload into disk I/Os. So far I've been able to get up to 180MB/s in VM and up to 220MB/s on real HW (on the SSD disk). I am still not sure why the overhead is so big but it is more than the network can handle so I do not care so much. The interesting thing is that I've found that the disk reads are not buffered in the VM unless I create and mount FS on the disk or I use the disks somehow. Simply put: Lets do dd to read disk for the first time (the /dev/vdd is an old Raptor disk 70MB/s is its real speed): [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 36.8038 s, 71.2 MB/s Buffers: 14444 kB Rereading the data shows that they are cached somewhere but not in buffers of the VM. Also the speed increased to "only" 500MB/s. The VM has 4GB of RAM (more that the test file) [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.16016 s, 508 MB/s Buffers: 14444 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.05727 s, 518 MB/s Buffers: 14444 kB Now lets mount the FS on /dev/vdd and try the dd again: [root@localhost ~]# mount /dev/vdd /mnt/tmp [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 4.68578 s, 559 MB/s Buffers: 2574592 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 1.50504 s, 1.7 GB/s Buffers: 2574592 kB While the first read was the same, all 2.6GB got buffered and the next read was at 1.7GB/s. And when I unmount the device: [root@localhost ~]# umount /mnt/tmp [root@localhost ~]# cat /proc/meminfo | grep Buffers Buffers: 14452 kB [root@localhost ~]# dd if=/dev/vdd of=/dev/null bs=256k count=10000 ; cat /proc/meminfo | grep Buffers 2621440000 bytes (2.6 GB) copied, 5.10499 s, 514 MB/s Buffers: 14468 kB The bcache was disabled while testing and the results are same on faster (newer) HDDs and on SSD (except for the initial read speed of course). To sum it up. When I read from the device via dd first time, it gets read from the disk. Next time I reread it gets cached in the host but not in the guest (thats actually the same issue, more on that later). When I mount the filesystem but try to read the device directly it gets cached in VM (via buffers). As soon as I stop "using" it, buffers are discarded and the device is not cached anymore in the VM. When I looked into buffers value on the host I realized that the situation is the same. The block I/O gets buffered only when the disk is in use, in this case it means "exported to a VM". On host, after all the measurement done: 3165552 buffers On the host, after the VM shutdown: 119176 buffers I know it is not important as the disks will be mounted all the time but I am curious and I would like to know why it is working like this.

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  • reiserfsck on lvm

    - by DaDaDom
    It seems like my filesystem got corrupted somehow during the last reboot of my server. I can't fsck some logical volumes anymore. The setup: root@rescue ~ # cat /mnt/rescue/etc/fstab proc /proc proc defaults 0 0 /dev/md0 /boot ext3 defaults 0 2 /dev/md1 / ext3 defaults,errors=remount-ro 0 1 /dev/systemlvm/home /home reiserfs defaults 0 0 /dev/systemlvm/usr /usr reiserfs defaults 0 0 /dev/systemlvm/var /var reiserfs defaults 0 0 /dev/systemlvm/tmp /tmp reiserfs noexec,nosuid 0 2 /dev/sda5 none swap defaults,pri=1 0 0 /dev/sdb5 none swap defaults,pri=1 0 0 [UPDATE] First question: what "part" should I check for bad blocks? The logical volume, the underlying /dev/md or the /dev/sdx below that? Is doing what I am doing the right way to go? [/UPDATE] The errormessage when checking /dev/systemlvm/usr: root@rescue ~ # reiserfsck /dev/systemlvm/usr reiserfsck 3.6.19 (2003 www.namesys.com) [...] Will read-only check consistency of the filesystem on /dev/systemlvm/usr Will put log info to 'stdout' Do you want to run this program?[N/Yes] (note need to type Yes if you do):Yes ########### reiserfsck --check started at Wed Feb 3 07:10:55 2010 ########### Replaying journal.. Reiserfs journal '/dev/systemlvm/usr' in blocks [18..8211]: 0 transactions replayed Checking internal tree.. Bad root block 0. (--rebuild-tree did not complete) Aborted Well so far, let's try --rebuild-tree: root@rescue ~ # reiserfsck --rebuild-tree /dev/systemlvm/usr reiserfsck 3.6.19 (2003 www.namesys.com) [...] Will rebuild the filesystem (/dev/systemlvm/usr) tree Will put log info to 'stdout' Do you want to run this program?[N/Yes] (note need to type Yes if you do):Yes Replaying journal.. Reiserfs journal '/dev/systemlvm/usr' in blocks [18..8211]: 0 transactions replayed ########### reiserfsck --rebuild-tree started at Wed Feb 3 07:12:27 2010 ########### Pass 0: ####### Pass 0 ####### Loading on-disk bitmap .. ok, 269716 blocks marked used Skipping 8250 blocks (super block, journal, bitmaps) 261466 blocks will be read 0%....20%....40%....60%....80%....100% left 0, 11368 /sec 52919 directory entries were hashed with "r5" hash. "r5" hash is selected Flushing..finished Read blocks (but not data blocks) 261466 Leaves among those 13086 Objectids found 53697 Pass 1 (will try to insert 13086 leaves): ####### Pass 1 ####### Looking for allocable blocks .. finished 0% left 12675, 0 /sec The problem has occurred looks like a hardware problem (perhaps memory). Send us the bug report only if the second run dies at the same place with the same block number. mark_block_used: (39508) used already Aborted Bad. But let's do it again as mentioned: [...] Flushing..finished Read blocks (but not data blocks) 261466 Leaves among those 13085 Objectids found 54305 Pass 1 (will try to insert 13085 leaves): ####### Pass 1 ####### Looking for allocable blocks .. finished 0%... left 12127, 958 /sec The problem has occurred looks like a hardware problem (perhaps memory). Send us the bug report only if the second run dies at the same place with the same block number. build_the_tree: Nothing but leaves are expected. Block 196736 - internal Aborted Same happens every time, only the actual error message changes. Sometimes I get mark_block_used: (somenumber) used already, other times the block number changes. Seems like something is REALLY broken. Are there any chances I can somehow get the partitions to work again? It's a server to which I don't have physical access directly (hosted server). Thanks in advance!

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