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  • RK4 Bouncing a Ball

    - by Jonathan Dickinson
    I am trying to wrap my head around RK4. I decided to do the most basic 'ball with gravity that bounces' simulation. I have implemented the following integrator given Glenn Fiedler's tutorial: /// <summary> /// Represents physics state. /// </summary> public struct State { // Also used internally as derivative. // S: Position // D: Velocity. /// <summary> /// Gets or sets the Position. /// </summary> public Vector2 X; // S: Position // D: Acceleration. /// <summary> /// Gets or sets the Velocity. /// </summary> public Vector2 V; } /// <summary> /// Calculates the force given the specified state. /// </summary> /// <param name="state">The state.</param> /// <param name="t">The time.</param> /// <param name="acceleration">The value that should be updated with the acceleration.</param> public delegate void EulerIntegrator(ref State state, float t, ref Vector2 acceleration); /// <summary> /// Represents the RK4 Integrator. /// </summary> public static class RK4 { private const float OneSixth = 1.0f / 6.0f; private static void Evaluate(EulerIntegrator integrator, ref State initial, float t, float dt, ref State derivative, ref State output) { var state = new State(); // These are a premature optimization. I like premature optimization. // So let's not concentrate on that. state.X.X = initial.X.X + derivative.X.X * dt; state.X.Y = initial.X.Y + derivative.X.Y * dt; state.V.X = initial.V.X + derivative.V.X * dt; state.V.Y = initial.V.Y + derivative.V.Y * dt; output = new State(); output.X.X = state.V.X; output.X.Y = state.V.Y; integrator(ref state, t + dt, ref output.V); } /// <summary> /// Performs RK4 integration over the specified state. /// </summary> /// <param name="eulerIntegrator">The euler integrator.</param> /// <param name="state">The state.</param> /// <param name="t">The t.</param> /// <param name="dt">The dt.</param> public static void Integrate(EulerIntegrator eulerIntegrator, ref State state, float t, float dt) { var a = new State(); var b = new State(); var c = new State(); var d = new State(); Evaluate(eulerIntegrator, ref state, t, 0.0f, ref a, ref a); Evaluate(eulerIntegrator, ref state, t + dt * 0.5f, dt * 0.5f, ref a, ref b); Evaluate(eulerIntegrator, ref state, t + dt * 0.5f, dt * 0.5f, ref b, ref c); Evaluate(eulerIntegrator, ref state, t + dt, dt, ref c, ref d); a.X.X = OneSixth * (a.X.X + 2.0f * (b.X.X + c.X.X) + d.X.X); a.X.Y = OneSixth * (a.X.Y + 2.0f * (b.X.Y + c.X.Y) + d.X.Y); a.V.X = OneSixth * (a.V.X + 2.0f * (b.V.X + c.V.X) + d.V.X); a.V.Y = OneSixth * (a.V.Y + 2.0f * (b.V.Y + c.V.Y) + d.V.Y); state.X.X = state.X.X + a.X.X * dt; state.X.Y = state.X.Y + a.X.Y * dt; state.V.X = state.V.X + a.V.X * dt; state.V.Y = state.V.Y + a.V.Y * dt; } } After reading over the tutorial I noticed a few things that just seemed 'out' to me. Notably how the entire simulation revolves around t at 0 and state at 0 - considering that we are working out a curve over the duration it seems logical that RK4 wouldn't be able to handle this simple scenario. Never-the-less I forged on and wrote a very simple Euler integrator: static void Integrator(ref State state, float t, ref Vector2 acceleration) { if (state.X.Y > 100 && state.V.Y > 0) { // Bounce vertically. acceleration.Y = -state.V.Y * t; } else { acceleration.Y = 9.8f; } } I then ran the code against a simple fixed-time step loop and this is what I got: 0.05 0.20 0.44 0.78 1.23 1.76 ... 74.53 78.40 82.37 86.44 90.60 94.86 99.23 103.05 105.45 106.94 107.86 108.42 108.76 108.96 109.08 109.15 109.19 109.21 109.23 109.23 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 109.24 ... As I said, I was expecting it to break - however I am unsure of how to fix it. I am currently looking into keeping the previous state and time, and working from that - although at the same time I assume that will defeat the purpose of RK4. How would I get this simulation to print the expected results?

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  • How Estimates Became Quotes

    - by Lee Brandt
    It’s our fault. Well, not completely, but we haven’t helped the situation any. All of what follows comes from my own experiences which, from talking to lots of other developers about it, seems to be pretty much par for the course. Where We Started When we first started estimating, we estimated pretty clearly. We would try to imagine something we’d done that was similar to the project being estimated and we’d toss it about in our heads a bit and see how much bigger or smaller we thought this new thing was, and add or subtract accordingly. We wouldn’t spend too much time on it, because we wanted to get to writing the software. Eventually, we’d come across some huge problem that there was now way we could’ve known about ahead of time. Either we didn’t see this thing or, we didn’t realize that this particular version of a problem would be so… problematic. We usually call this “not knowing what we don’t know”. It’s unavoidable. We just can’t know. Until we wade in and start putting some code together, there are just some things we won’t know… and some things we don’t even know that we don’t know. Y’know? So what happens? We go over budget. Project managers scream and dance the dance of the stressed-out project manager, and there is nothing we can do (or could’ve done) about it. We didn’t know. We thought about it for a bit and we didn’t see this herculean task sitting in the middle of our nice quiet project, and it has bitten us in the rear end. We now know how to handle this in the future, though. We will take some more time to pick around the requirements and discover all those things we don’t know. We’ll do some prototyping, we’ll read some blogs about similar projects, we’ll really grill the customer with questions during the requirements gathering phase. We’ll keeping asking “what else?” until the shove us down the stairs. We’ll take our time and uncover it all. We Learned, But Good The next time comes, and you know what happens? We do it. We grill the customer for weeks and prototype and read and research and we estimate everything down to the last button on the last form. Know what that gets us? It gets us three months of wasted time, and our estimate will still be off. Possibly off by a factor of four. WTF, mate? No way we could be surprised by something! We uncovered every particle. We turned every stone. How is it we still came across unknowns? Because we STILL didn’t know what we didn’t know. How could we? We didn’t know to ask. The worst part is, we’ve now convinced the product that this is NOT an estimate. It is a solid number based on massive research and an endless number of questions that they answered. There is absolutely now way you don’t know everything there is to know about this project now. No way there is anything you haven’t uncovered. And their faith in that “Esti-Quote” goes through the roof. When the project goes over this time, they might even begin to question whether or not you know what you’re doing. Who could blame them? You drilled them for weeks about every little thing, and when they complained about all the questions, you told them you wanted to uncover everything so there would be no surprises. SO we set them up to faile Guess, Then Plan We had a chance. At the beginning we could have just said, “That’s just a gut-feeling estimate, based on my past experience with similar projects. There could still be surprises.” If we spend SOME time doing SOME discovery and then bounce that against our own past experiences, we can come up with a fairly healthy estimate. We can then help the product owner understand that an estimate is a guess. Sure, it’s an educated guess, but it is still a guess. If we get it right it will be almost completely luck. Then, we help them to plan the development by taking that guess (yes, they still need the guess for planning purposes) and start measuring early and often to see if we still think we are right. We should adjust the estimate and alert the product owner as soon as we see problems (bad news does not age well) and we should be able to see any problems immediately if we are constantly measuring our pace. In lean software, we start with that guess and begin measuring cycle times immediately. Then we can make projections based on those cycle times and compare them to our estimate. This constant feedback is the best way to ensure that there are no surprises at the END of the project. There sill still be surprises, but we’ll see them sooner and have a better understanding of how they will affect our overall timeline. What do you think?

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  • How do I cleanly design a central render/animation loop?

    - by mtoast
    I'm learning some graphics programming, and am in the midst of my first such project of any substance. But, I am really struggling at the moment with how to architect it cleanly. Let me explain. To display complicated graphics in my current language of choice (JavaScript -- have you heard of it?), you have to draw graphical content onto a <canvas> element. And to do animation, you must clear the <canvas> after every frame (unless you want previous graphics to remain). Thus, most canvas-related JavaScript demos I've seen have a function like this: function render() { clearCanvas(); // draw stuff here requestAnimationFrame(render); } render, as you may surmise, encapsulates the drawing of a single frame. What a single frame contains at a specific point in time, well... that is determined by the program state. So, in order for my program to do its thing, I just need to look at the state, and decide what to render. Right? Right. But that is more complicated than it seems. My program is called "Critter Clicker". In my program, you see several cute critters bouncing around the screen. Clicking on one of them agitates it, making it bounce around even more. There is also a start screen, which says "Click to start!" prior to the critters being displayed. Here are a few of the objects I'm working with in my program: StartScreenView // represents the start screen CritterTubView // represents the area in which the critters live CritterList // a collection of all the critters Critter // a single critter model CritterView // view of a single critter Nothing too egregious with this, I think. Yet, when I set out to flesh out my render function, I get stuck, because everything I write seems utterly ugly and reminiscent of a certain popular Italian dish. Here are a couple of approaches I've attempted, with my internal thought process included, and unrelated bits excluded for clarity. Approach 1: "It's conditions all the way down" // "I'll just write the program as I think it, one frame at a time." if (assetsLoaded) { if (userClickedToStart) { if (critterTubDisplayed) { if (crittersDisplayed) { forEach(crittersList, function(c) { if (c.wasClickedRecently) { c.getAgitated(); } }); } else { displayCritters(); } } else { displayCritterTub(); } } else { displayStartScreen(); } } That's a very much simplified example. Yet even with only a fraction of all the rendering conditions visible, render is already starting to get out of hand. So, I dispense with that and try another idea: Approach 2: Under the Rug // "Each view object shall be responsible for its own rendering. // "I'll pass each object the program state, and each can render itself." startScreen.render(state); critterTub.render(state); critterList.render(state); In this setup, I've essentially just pushed those crazy nested conditions to a deeper level in the code, hiding them from view. In other words, startScreen.render would check state to see if it needed actually to be drawn or not, and take the correct action. But this seems more like it only solves a code-aesthetic problem. The third and final approach I'm considering that I'll share is the idea that I could invent my own "wheel" to take care of this. I'm envisioning a function that takes a data structure that defines what should happen at any given point in the render call -- revealing the conditions and dependencies as a kind of tree. Approach 3: Mad Scientist renderTree({ phases: ['startScreen', 'critterTub', 'endCredits'], dependencies: { startScreen: ['assetsLoaded'], critterTub: ['startScreenClicked'], critterList ['critterTubDisplayed'] // etc. }, exclusions: { startScreen: ['startScreenClicked'], // etc. } }); That seems kind of cool. I'm not exactly sure how it would actually work, but I can see it being a rather nifty way to express things, especially if I flex some of JavaScript's events. In any case, I'm a little bit stumped because I don't see an obvious way to do this. If you couldn't tell, I'm coming to this from the web development world, and finding that doing animation is a bit more exotic than arranging an MVC application for handling simple requests - responses. What is the clean, established solution to this common-I-would-think problem?

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  • UIScrollView image/photo viewer with paging enabled and zooming

    - by Mike Weller
    OK, I think it's time to make an official place on the internet for this problem: How to make a UIScrollView photoviewer with paging and zooming. Welcome my fellow UIScrollView hackers. I have a UIScrollView with paging enabled, and I'm displaying UIImageViews like the built-in photos app. (Does this sound familiar yet?) I found the following project on github: http://wiki.github.com/andreyvit/ScrollingMadness Which shows how to implement zooming in a scroll view while paging is enabled. If anyone else tries this out, I actually had to remove the UIScrollView subclass and use the native class otherwise it doesn't work. I think it's because of changes in the 3.0 SDK relating to how the scroll view intercepts touch events. So the the idea is to remove all the other views when you start zooming, and move the current view to (0, 0) in the scrollview, updating the contentsize etc. Then when you zoom back to 1.0f it adds the other views back and puts things all back in order. Anyway, that project works perfectly in the simulator, but on the device there is some nasty movement of the view you are resizing, which looks like it's caused by the fact we are changing the contentsize/offset etc. for the view being resized. You have to do this view moving otherwise you can pan left through the whitespace left by the other views. I found one interesting note in the "Known Issues" of the 3.0 SDK release notes: UIScrollView: After zooming, content inset is ignored and content is left in the wrong position. This kind of sounds like what is happening here. After zooming in, the view will shift offscreen because you have changed the offset etc. I've spent hours on this already and I'm slowing coming to the sad realization that this just isn't going to work. Three20's photo viewer is out of the question: it's too heavy weight and there is too much unnecessary UI and other behaviour. The built in Photo app seems to do some magic. If you zoom in on an image and pan to the far edges, the current photo moves independently of the photo next to it which isn't what you get when trying this with a standard UIScrollView. I've seen discussion about nesting the UIScrollView's but I really don't want to go there. Has anybody managed this with the standard UIScrollView (and works in the 2.2 and 3.0 SDK)? I don't fancy rolling my own zoom + bounce + pan + paging code.

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  • BounceEase and silverlight 4 BarSeries

    - by Pharabus
    Hi, I am trying to get a bar series to "bounce" when drawing, I assumed the BounceEase TransitionEasingFunction would do this but the lines just fade in, I have posted the xaml and code behind below, does anyone know where I have gone wrong or is it more complex than I though, I am fairly new to silverlight XAML <Grid x:Name="LayoutRoot" Background="White"> <chartingToolkit:Chart x:Name="MyChart"> <chartingToolkit:BarSeries Title="Sales" ItemsSource="{Binding}" IndependentValuePath="Name" DependentValuePath="Value" AnimationSequence="FirstToLast" TransitionDuration="00:00:3"> <chartingToolkit:BarSeries.TransitionEasingFunction> <BounceEase EasingMode="EaseInOut" Bounciness="5" /> </chartingToolkit:BarSeries.TransitionEasingFunction> <chartingToolkit:BarSeries.DataPointStyle> <Style TargetType="Control"> <Setter Property="Background" Value="Red"/> </Style> </chartingToolkit:BarSeries.DataPointStyle> </chartingToolkit:BarSeries> <chartingToolkit:Chart.Axes> <chartingToolkit:LinearAxis Title="Types owned" Orientation="X" Minimum="0" Maximum="300" Interval="10" ShowGridLines="True" FontStyle='Italic'/> </chartingToolkit:Chart.Axes> </chartingToolkit:Chart> </Grid> code behind public class MyClass : DependencyObject { public string Name { get; set; } public Double Value { get { return (Double)GetValue(myValueProperty); } set{SetValue(myValueProperty,value);} } public static readonly DependencyProperty myValueProperty = DependencyProperty.Register("Value", typeof(Double), typeof(MyClass), null); } public MainPage() { InitializeComponent(); //Get the data IList<MyClass> l = this.GetData(); //Get a reference to the SL Chart MyChart.DataContext = l.OrderBy(e => e.Value); //Find the highest number and round it up to the next digit DispatcherTimer myDispatcherTimer = new DispatcherTimer(); myDispatcherTimer.Interval = new TimeSpan(0, 0, 0, 5, 0); // 100 Milliseconds myDispatcherTimer.Tick += new EventHandler(Each_Tick); myDispatcherTimer.Start(); } public void Each_Tick(object o, EventArgs sender) { ((BarSeries)MyChart.Series[0]).DataContext = GetData(); } private IList<MyClass> GetData() { Random random = new Random(); return new List<MyClass>() { new MyClass() {Name="Bob Zero",Value=(random.NextDouble() * 100.0)}, new MyClass() {Name="Bob One",Value=(random.NextDouble() * 100.0)}, new MyClass() {Name="Bob Two",Value=(random.NextDouble() * 100.0)}, new MyClass() {Name="Bob Three",Value=(random.NextDouble() * 100.0)} }; }

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  • SQL Server to PostgreSQL - Migration and design concerns

    - by youwhut
    Currently migrating from SQL Server to PostgreSQL and attempting to improve a couple of key areas on the way: I have an Articles table: CREATE TABLE [dbo].[Articles]( [server_ref] [int] NOT NULL, [article_ref] [int] NOT NULL, [article_title] [varchar](400) NOT NULL, [category_ref] [int] NOT NULL, [size] [bigint] NOT NULL ) Data (comma delimited text files) is dumped on the import server by ~500 (out of ~1000) servers on a daily basis. Importing: Indexes are disabled on the Articles table. For each dumped text file Data is BULK copied to a temporary table. Temporary table is updated. Old data for the server is dropped from the Articles table. Temporary table data is copied to Articles table. Temporary table dropped. Once this process is complete for all servers the indexes are built and the new database is copied to a web server. I am reasonably happy with this process but there is always room for improvement as I strive for a real-time (haha!) system. Is what I am doing correct? The Articles table contains ~500 million records and is expected to grow. Searching across this table is okay but could be better. i.e. SELECT * FROM Articles WHERE server_ref=33 AND article_title LIKE '%criteria%' has been satisfactory but I want to improve the speed of searching. Obviously the "LIKE" is my problem here. Suggestions? SELECT * FROM Articles WHERE article_title LIKE '%criteria%' is horrendous. Partitioning is a feature of SQL Server Enterprise but $$$ which is one of the many exciting prospects of PostgreSQL. What performance hit will be incurred for the import process (drop data, insert data) and building indexes? Will the database grow by a huge amount? The database currently stands at 200 GB and will grow. Copying this across the network is not ideal but it works. I am putting thought into changing the hardware structure of the system. The thought process of having an import server and a web server is so that the import server can do the dirty work (WITHOUT indexes) while the web server (WITH indexes) can present reports. Maybe reducing the system down to one server would work to skip the copying across the network stage. This one server would have two versions of the database: one with the indexes for delivering reports and the other without for importing new data. The databases would swap daily. Thoughts? This is a fantastic system, and believe it or not there is some method to my madness by giving it a big shake up. UPDATE: I am not looking for help with relational databases, but hoping to bounce ideas around with data warehouse experts.

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  • NSStringWithFormat Swizzled to allow missing format numbered args

    - by coneybeare
    Based on this SO question asked a few hours ago, I have decided to implement a swizzled method that will allow me to take a formatted NSString as the format arg into stringWithFormat, and have it not break when omitting one of the numbered arg references (%1$@, %2$@) I have it working, but this is the first copy, and seeing as this method is going to be potentially called hundreds of thousands of times per app run, I need to bounce this off of some experts to see if this method has any red flags, major performance hits, or optimizations #define NUMARGS(...) (sizeof((int[]){__VA_ARGS__})/sizeof(int)) @implementation NSString (UAFormatOmissions) + (id)uaStringWithFormat:(NSString *)format, ... { if (format != nil) { va_list args; va_start(args, format); // $@ is an ordered variable (%1$@, %2$@...) if ([format rangeOfString:@"$@"].location == NSNotFound) { //call apples method NSString *s = [[[NSString alloc] initWithFormat:format arguments:args] autorelease]; va_end(args); return s; } NSMutableArray *newArgs = (NSMutableArray *)[NSMutableArray arrayWithCapacity:NUMARGS(args)]; id arg = nil; int i = 1; while (arg = va_arg(args, id)) { NSString *f = (NSString *)[NSString stringWithFormat:@"%%%d\$\@", i]; i++; if ([format rangeOfString:f].location == NSNotFound) continue; else [newArgs addObject:arg]; } va_end(args); char *newArgList = (char *)malloc(sizeof(id) * [newArgs count]); [newArgs getObjects:(id *)newArgList]; NSString* result = [[[NSString alloc] initWithFormat:format arguments:newArgList] autorelease]; free(newArgList); return result; } return nil; } The basic algorithm is: search the format string for the %1$@, %2$@ variables by searching for %@ if not found, call the normal stringWithFormat and return else, loop over the args if the format has a position variable (%i$@) for position i, add the arg to the new arg array else, don't add the arg take the new arg array, convert it back into a va_list, and call initWithFormat:arguments: to get the correct string. The idea is that I would run all [NSString stringWithFormat:] calls through this method instead. This might seem unnecessary to many, but click on to the referenced SO question (first line) to see examples of why I need to do this. Ideas? Thoughts? Better implementations? Better Solutions?

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  • NSString stringWithFormat swizzled to allow missing format numbered args

    - by coneybeare
    Based on this SO question asked a few hours ago, I have decided to implement a swizzled method that will allow me to take a formatted NSString as the format arg into stringWithFormat, and have it not break when omitting one of the numbered arg references (%1$@, %2$@) I have it working, but this is the first copy, and seeing as this method is going to be potentially called hundreds of thousands of times per app run, I need to bounce this off of some experts to see if this method has any red flags, major performance hits, or optimizations #define NUMARGS(...) (sizeof((int[]){__VA_ARGS__})/sizeof(int)) @implementation NSString (UAFormatOmissions) + (id)uaStringWithFormat:(NSString *)format, ... { if (format != nil) { va_list args; va_start(args, format); // $@ is an ordered variable (%1$@, %2$@...) if ([format rangeOfString:@"$@"].location == NSNotFound) { //call apples method NSString *s = [[[NSString alloc] initWithFormat:format arguments:args] autorelease]; va_end(args); return s; } NSMutableArray *newArgs = (NSMutableArray *)[NSMutableArray arrayWithCapacity:NUMARGS(args)]; id arg = nil; int i = 1; while (arg = va_arg(args, id)) { NSString *f = (NSString *)[NSString stringWithFormat:@"%%%d\$\@", i]; i++; if ([format rangeOfString:f].location == NSNotFound) continue; else [newArgs addObject:arg]; } va_end(args); char *newArgList = (char *)malloc(sizeof(id) * [newArgs count]); [newArgs getObjects:(id *)newArgList]; NSString* result = [[[NSString alloc] initWithFormat:format arguments:newArgList] autorelease]; free(newArgList); return result; } return nil; } The basic algorithm is: search the format string for the %1$@, %2$@ variables by searching for %@ if not found, call the normal stringWithFormat and return else, loop over the args if the format has a position variable (%i$@) for position i, add the arg to the new arg array else, don't add the arg take the new arg array, convert it back into a va_list, and call initWithFormat:arguments: to get the correct string. The idea is that I would run all [NSString stringWithFormat:] calls through this method instead. This might seem unnecessary to many, but click on to the referenced SO question (first line) to see examples of why I need to do this. Ideas? Thoughts? Better implementations? Better Solutions?

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  • Potential issues using member's "from" address and the "sender" header

    - by Paul Burney
    Hi all, A major component of our application sends email to members on behalf of other members. Currently we set the "From" address to our system address and use a "Reply-to" header with the member's address. The issue is that replies from some email clients (and auto-replies/bounces) don't respect the "Reply-to" header so get sent to our system address, effectively sending them to a black hole. We're considering setting the "From" address to our member's address, and the "Sender" address to our system address. It appears this way would pass SPF and Sender-ID checks. Are there any reasons not to switch to this method? Are there any other potential issues? Thanks in advance, -Paul Here are way more details than you probably need: When the application was first developed, we just changed the "from" address to be that of the sending member as that was the common practice at the time (this was many years ago). We later changed that to have the "from" address be the member's name and our address, i.e., From: "Mary Smith" <[email protected]> With a "reply-to" header set to the member's address: Reply-To: "Mary Smith" <[email protected]> This helped with messages being mis-categorized as spam. As SPF became more popular, we added an additional header that would work in conjunction with our SPF records: Sender: <[email protected]> Things work OK, but it turns out that, in practice, some email clients and most MTA's don't respect the "Reply-To" header. Because of this, many members send messages to [email protected] instead of the desired member. So, I started envisioning various schemes to add data about the sender to the email headers or encode it in the "from" email address so that we could process the response and redirect appropriately. For example, From: "Mary Smith" <[email protected]> where the string after "messages" is a hash representing Mary Smith's member in our system. Of course, that path could lead to a lot of pain as we need to develop MTA functionality for our system address. I was looking again at the SPF documentation and found this page interesting: http://www.openspf.org/Best_Practices/Webgenerated They show two examples, that of evite.com and that of egreetings.com. Basically, evite.com is doing it the way we're doing it. The egreetings.com example uses the member's from address with an added "Sender" header. So the question is, are there any potential issues with using the egreetings method of the member's from address with a sender header? That would eliminate the replies that bad clients send to the system address. I don't believe that it solves the bounce/vacation/whitelist issue since those often send to the MAIL FROM even if Return Path is specified.

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  • How to implement code for multiple buttons using c++ in Silverlight for Windows Embedded

    - by Abhi
    Dear all I have referred the following link: Silverlight for Windows Embedded By referring this link i created a demo application which consist of two buttons created using Microsoft expression blend 2 tools. And then written a code referring the above site. Now my button names are "Browser Button" and "Media Button". On click of any one of the button i should able to launch the respective application. I was able to do for "Browser Button" but not for "Media Button" and if i do for "Media Button" then i am not able to do for "Browser Button".. I mean to say that how should i create event handler for both the buttons. This is the code in c++ which i should modify class BtnEventHandler { public: HRESULT OnClick(IXRDependencyObject* source,XRMouseButtonEventArgs* args) { RETAILMSG(1,(L"Browser event")); Execute(L"\\Windows\\iesample.exe",L""); return S_OK; } }; // entry point for the application. INT WINAPI WinMain(HINSTANCE hInstance,HINSTANCE hPrevInstance, LPWSTR lpCmdLine,int nCmdShow) { PrintMessage(); int exitCode = -1; HRESULT hr = S_OK; if (!XamlRuntimeInitialize()) return -1; HRESULT retcode; IXRApplicationPtr app; if (FAILED(retcode=GetXRApplicationInstance(&app))) return -1; if (FAILED(retcode=app->AddResourceModule(hInstance))) return -1; XRWindowCreateParams wp; ZeroMemory(&wp, sizeof(XRWindowCreateParams)); wp.Style = WS_OVERLAPPED; wp.pTitle = L"Bounce Test"; wp.Left = 0; wp.Top = 0; XRXamlSource xamlsrc; xamlsrc.SetResource(hInstance,TEXT("XAML"),MAKEINTRESOURCE(IDR_XAML1)); IXRVisualHostPtr vhost; if (FAILED(retcode=app->CreateHostFromXaml(&xamlsrc, &wp, &vhost))) return -1; IXRFrameworkElementPtr root; if (FAILED(retcode=vhost->GetRootElement(&root))) return -1; IXRButtonBasePtr btn; if (FAILED(retcode=root->FindName(TEXT("BrowserButton"), &btn))) return -1; IXRDelegate<XRMouseButtonEventArgs>* clickdelegate; BtnEventHandler handler; if(FAILED(retcode=CreateDelegate (&handler,&BtnEventHandler::OnClick,&clickdelegate))) return -1; if (FAILED(retcode=btn->AddClickEventHandler(clickdelegate))) return -1; UINT exitcode; if (FAILED(retcode=vhost->StartDialog(&exitcode))) return -1; return exitCode; } I have to add event handler for both the button so that on emulator whenever i click on any one of the button i should be able to launch the respective applications. Thanks in advance

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  • ArrayAdapter throwing ArrayIndexOutOfBoundsException

    - by alex
    I am getting an error of an array out of bounce error when i am using my custom array adapter. I am wondering if there are any coding errors I have overlooked. Here is the error log 06-10 20:21:53.254: E/AndroidRuntime(315): FATAL EXCEPTION: main 06-10 20:21:53.254: E/AndroidRuntime(315): java.lang.RuntimeException: Unable to start activity ComponentInfo{alex.android.galaxy.tab.latest/alex.android.galaxy.tab.latest.Basic_db_output}: java.lang.ArrayIndexOutOfBoundsException 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread.performLaunchActivity(ActivityThread.java:2663) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread.handleLaunchActivity(ActivityThread.java:2679) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread.access$2300(ActivityThread.java:125) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread$H.handleMessage(ActivityThread.java:2033) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.os.Handler.dispatchMessage(Handler.java:99) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.os.Looper.loop(Looper.java:123) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread.main(ActivityThread.java:4627) 06-10 20:21:53.254: E/AndroidRuntime(315): at java.lang.reflect.Method.invokeNative(Native Method) 06-10 20:21:53.254: E/AndroidRuntime(315): at java.lang.reflect.Method.invoke(Method.java:521) 06-10 20:21:53.254: E/AndroidRuntime(315): at com.android.internal.os.ZygoteInit$MethodAndArgsCaller.run(ZygoteInit.java:868) 06-10 20:21:53.254: E/AndroidRuntime(315): at com.android.internal.os.ZygoteInit.main(ZygoteInit.java:626) 06-10 20:21:53.254: E/AndroidRuntime(315): at dalvik.system.NativeStart.main(Native Method) 06-10 20:21:53.254: E/AndroidRuntime(315): Caused by: java.lang.ArrayIndexOutOfBoundsException 06-10 20:21:53.254: E/AndroidRuntime(315): at alex.android.galaxy.tab.latest.Basic_db_output.onCreate(Basic_db_output.java:44) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.Instrumentation.callActivityOnCreate(Instrumentation.java:1047) 06-10 20:21:53.254: E/AndroidRuntime(315): at android.app.ActivityThread.performLaunchActivity(ActivityThread.java:2627) I am basing my code example on this How to use ArrayAdapter<myClass> ArrayList list = new ArrayList(); for(int i=1; i <= 3; i++) { Reader reader = new ResultsReader("android_galaxy_tab_latest/src/quiz"+i+".txt"); reader.read(); String str = ((ResultsReader)reader).getInput(); String data[] = str.split("<.>"); Question q = new Question(); q.question = data[0]; q.answer = Integer.parseInt(data[1]); q.choice1 = data[2]; q.choice2 = data[3]; q.choice3 = data[4]; list.add(q); }

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  • proper fill an image larger than screen

    - by madcat
    what I wanted to achieve here is simply fit the image width to the screen on both orientations and use UIScrollView to just allow scroll vertically to see the whole image. both viewController and view are created pragmatically. the image loaded is larger than screen on both width and height. here is the related code in my viewController: - (BOOL)shouldAutorotateToInterfaceOrientation:(UIInterfaceOrientation)interfaceOrientation { return YES; } - (void)loadView { UIScreen *screen = [UIScreen mainScreen]; CGRect rect = [screen applicationFrame]; self.view = [[UIView alloc] initWithFrame:rect]; self.view.contentMode = UIViewContentModeScaleAspectFill; self.view.autoresizingMask = UIViewAutoresizingFlexibleWidth | UIViewAutoresizingFlexibleHeight; UIImage *img=[[UIImage alloc] initWithContentsOfFile:[[NSBundle mainBundle] pathForResource:@"image" ofType:@"png"]]; UIImageView *imgView =[[UIImageView alloc] initWithImage:img]; [img release]; imgView.contentMode = UIViewContentModeScaleAspectFill; imgView.autoresizingMask = UIViewAutoresizingFlexibleWidth | UIViewAutoresizingFlexibleHeight; [self.view addSubview:imgView]; [imgView release]; } tried all combinations for both contentMode above, did not give me correct result. the most close I am getting now: I manually resize imgView in loadView, portrait mode would display correctly since app always starts with portrait mode, but in landscape mode, the width fits correctly, but image is centered vertically rather than top aligned. if I add the imgView to a scrollView, in landscape mode it looks like contentSize is not set to full image size. but when I scroll bounce I can see the image is there in full size. question: why I need to resize it manually? in landscape mode how and where I can 'move' the imgView, so imgView.frame.origin is (0,0) and works correctly with a scroll view? Thanks! UPDATE: I added: imgView.clipsToBounds = YES; and find out in landscape mode the image bounds is smaller than screen in height. so the question becomes how to have the image view keeps original ratio (thus shows the full image always) when rotated to landscape? do I need to manually resize it after rotation again?

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  • PHP send batch email [closed]

    - by qalbiol
    Possible Duplicate: Sending mass email using PHP I have a PHP script that sends an individual email to all users in my DB, such as a monthly / weekly newsletter. The code I am using goes as follows: $subject = $_POST['subject']; $message = $_POST['message']; // Get all the mailing list subscribers. $query = $db->prepare("SELECT * FROM maildb"); $query->execute(); // Loop through all susbcribers, and send and individual email. foreach ($query as $row) { // Setting maximum time limit to infinite. set_time_limit(0); $newMessage = '<!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head> </head> <body>'; // Search for the [unsubscribe] tag and replace it with a URL for the user to unsubscribe $newMessage .= str_replace("[unsubscribe]", "<a href='".BASE_URL."unsubscribe/".$row['hash']."/".$row['email']."'>unsubscribe</a>", $message); $newMessage .= '</body></html>'; $to = $row['email']; // Establish content headers $headers = "From: [email protected]"."\n"; $headers .= "Reply-To: [email protected]"."\n"; $headers .= "X-Mailer: PHP v.". phpversion()."\n"; $headers .= "MIME-Version: 1.0"."\n"; $headers .= "Content-Type: text/html; charset=iso-8859-1"."\n"; $headers .= "Content-Transfer-Encoding: 8bit;"; mail($to, $subject, $newMessage, $headers); // Send email to each individual user } This code works perfectly with a REALLY small database... I recently populated my test db with 200k+ users, and obviously this script fails, gets out of memory, and dies... I know this is a bad way to send so many emails, thats why I'd like to ask you for much more efficient ways to do this! Thank you very much!

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  • Splitting a raidctl mirror safely

    - by milkfilk
    I have a Sun T5220 server with the onboard LSI card and two disks that were in a RAID 1 mirror. The data is not important right now but we had a failed disk and are trying to understand how to do this for real if we had to recover from a failure. The initial situation looked like this: # raidctl -l c1t0d0 Volume Size Stripe Status Cache RAID Sub Size Level Disk ---------------------------------------------------------------- c1t0d0 136.6G N/A DEGRADED OFF RAID1 0.1.0 136.6G GOOD N/A 136.6G FAILED Green light on the 0.0.0 disk. Find / lights up the 0.1.0 disk. So I know I have a bad drive and which one it is. Server still boots obviously. First, we tried putting a new disk in. This disk came from an unknown source. Format would not see it, cfgadm -al would not see it so raidctl -l would not see it. I figure it's bad. We tried another disk from another spare server: # raidctl -c c1t1d0 c1t0d0 (where t1 is my good disk - 0.1.0) Disk has occupied space. Also the different syntax options don't change anything: # raidctl -C "0.1.0 0.0.0" -r 1 1 Disk has occupied space. # raidctl -C "0.1.0 0.0.0" 1 Disk has occupied space. Ok. Maybe this is because the disk from the spare server had a RAID 1 on it already. Aha, I can see another volume in raidctl: # raidctl -l Controller: 1 Volume:c1t1d0 (this is my server's root mirror) Volume:c1t132d0 (this is the foreign root mirror) Disk: 0.0.0 Disk: 0.1.0 ... No problem. I don't care about the data, I'll just delete the foreign mirror. # raidctl -d c1t132d0 (warning about data deletion but it works) At this point, /usr/bin/ binaries freak out. By that I mean, ls -l /usr/bin/which shows 1.4k but cat /usr/bin/which gives me a newline. Great, I just blew away the binaries (ie: binaries in mem still work)? I bounce the box. It all comes back fine. WTF. Anyway, back to recreating my mirror. # raidctl -l Controller: 1 Volume:c1t1d0 (this is my server's root mirror) Disk: 0.0.0 Disk: 0.1.0 ... Man says that you can delete a mirror and it will split it. Ok, I'll delete the root mirror. # raidctl -d c1t0d0 Array in use. (this might not be the exact error) I googled this and found of course you can't do this (even with -f) while booted off the mirror. Ok. I boot cdrom -s and deleted the volume. Now I have one disk that has a type of "LSI-Logical-Volume" on c1t1d0 (where my data is) and a brand new "Hitachi 146GB" on c1t0d0 (what I'm trying to mirror to): (booted off the CD) # raidctl -c c1t1d0 c1t0d0 (man says it's source destination for mirroring) Illegal Array Layout. # raidctl -C "0.1.0 0.0.0" -r 1 1 (alt syntax per man) Illegal Array Layout. # raidctl -C "0.1.0 0.0.0" 1 (assumes raid1, no help) Illegal Array Layout. Same size disks, same manufacturer but I did delete the volume instead of throwing in a blank disk and waiting for it to resync. Maybe this was a critical error. I tried selecting the type in format for my good disk to be a plain 146gb disk but it resets the partition table which I'm pretty sure would wipe the data (bad if this was production). Am I boned? Anyone have experience with breaking and resyncing a mirror? There's nothing on Google about "Illegal Array Layout" so here's my contrib to the search gods.

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  • Postfix - Gmail - Mountain Lion // can't send mail

    - by miako
    I have read most of the tutorials found on google but still can't make it work. I run the command : date | mail -s "Test" [email protected] . The log is this : Oct 22 11:38:00 XXX.local postfix/master[288]: daemon started -- version 2.9.2, configuration /etc/postfix Oct 22 11:38:00 XXX.local postfix/pickup[289]: 9D85418A031: uid=501 from=<me> Oct 22 11:38:00 XXX.local postfix/cleanup[291]: 9D85418A031: message-id=<[email protected]> Oct 22 11:38:00 XXX.local postfix/qmgr[290]: 9D85418A031: from=<[email protected]>, size=327, nrcpt=1 (queue active) Oct 22 11:38:00 XXX.local postfix/smtp[293]: initializing the client-side TLS engine Oct 22 11:38:02 XXX.local postfix/smtp[293]: setting up TLS connection to smtp.gmail.com[173.194.70.109]:587 Oct 22 11:38:02 XXX.local postfix/smtp[293]: smtp.gmail.com[173.194.70.109]:587: TLS cipher list "ALL:!EXPORT:!LOW:+RC4:@STRENGTH:!eNULL" Oct 22 11:38:02 XXX.local postfix/smtp[293]: SSL_connect:before/connect initialization Oct 22 11:38:02 XXX.local postfix/smtp[293]: SSL_connect:SSLv2/v3 write client hello A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 read server hello A Oct 22 11:38:03 XXX.local postfix/smtp[293]: smtp.gmail.com[173.194.70.109]:587: certificate verification depth=2 verify=0 subject=/C=US/O=GeoTrust Inc./CN=GeoTrust Global CA Oct 22 11:38:03 --- last message repeated 1 time --- Oct 22 11:38:03 XXX.local postfix/smtp[293]: smtp.gmail.com[173.194.70.109]:587: certificate verification depth=1 verify=1 subject=/C=US/O=Google Inc/CN=Google Internet Authority G2 Oct 22 11:38:03 XXX.local postfix/smtp[293]: smtp.gmail.com[173.194.70.109]:587: certificate verification depth=0 verify=1 subject=/C=US/ST=California/L=Mountain View/O=Google Inc/CN=smtp.gmail.com Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 read server certificate A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 read server done A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 write client key exchange A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 write change cipher spec A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 write finished A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 flush data Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 read server session ticket A Oct 22 11:38:03 XXX.local postfix/smtp[293]: SSL_connect:SSLv3 read finished A Oct 22 11:38:03 XXX.local postfix/smtp[293]: smtp.gmail.com[173.194.70.109]:587: subject_CN=smtp.gmail.com, issuer_CN=Google Internet Authority G2, fingerprint E4:CA:10:85:C3:53:00:E6:A1:D2:AC:C4:35:E4:A2:10, pkey_fingerprint=D6:06:2E:15:AF:DF:E9:50:A5:B4:E2:E4:C5:2E:F9:BA Oct 22 11:38:03 XXX.local postfix/smtp[293]: Untrusted TLS connection established to smtp.gmail.com[173.194.70.109]:587: TLSv1 with cipher RC4-SHA (128/128 bits) Oct 22 11:38:03 XXX.local postfix/smtp[293]: 9D85418A031: to=<[email protected]>, relay=smtp.gmail.com[173.194.70.109]:587, delay=3.4, delays=0.26/0.13/2.8/0.26, dsn=5.5.1, status=bounced (host smtp.gmail.com[173.194.70.109] said: 530-5.5.1 Authentication Required. Learn more at 530 5.5.1 http://support.google.com/mail/bin/answer.py?answer=14257 s3sm54097220eeo.3 - gsmtp (in reply to MAIL FROM command)) Oct 22 11:38:04 XXX.local postfix/cleanup[291]: D4D2F18A03C: message-id=<[email protected]> Oct 22 11:38:04 XXX.local postfix/qmgr[290]: D4D2F18A03C: from=<>, size=2382, nrcpt=1 (queue active) Oct 22 11:38:04 XXX.local postfix/bounce[297]: 9D85418A031: sender non-delivery notification: D4D2F18A03C Oct 22 11:38:04 XXX.local postfix/qmgr[290]: 9D85418A031: removed Oct 22 11:38:04 XXX.local postfix/local[298]: D4D2F18A03C: to=<[email protected]>, relay=local, delay=0.11, delays=0/0.08/0/0.02, dsn=2.0.0, status=sent (delivered to mailbox) Oct 22 11:38:04 XXX.local postfix/qmgr[290]: D4D2F18A03C: removed Oct 22 11:39:00 XXX.local postfix/master[288]: master exit time has arrived I am really confused as i have never setup MTA again an i need it for local web development. I don't use XAMPP. I use the built in Servers. Can anyone guide me?

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  • Detecting source of memory usage on a Linux box

    - by apeace
    I have a toy Linux box with 256mb RAM running Ubuntu 10.04.1 LTS. Here is the output of free -m: total used free shared buffers cached Mem: 245 122 122 0 19 64 -/+ buffers/cache: 38 206 Swap: 511 0 511 Unless I'm reading this wrong, 122mb is being used and only 84mb of that is disk cache. Here are all processes I'm running sorted by memory usage (ps -eo pmem,pcpu,rss,vsize,args | sort -k 1 -r): %MEM %CPU RSS VSZ COMMAND 5.0 0.0 12648 633140 node /home/node/main/sites.js 1.5 0.0 3884 251736 /usr/sbin/console-kit-daemon --no-daemon 1.3 0.0 3328 77108 sshd: apeace [priv] 0.9 0.0 2344 19624 -bash 0.7 0.0 1776 23620 /sbin/init 0.6 0.0 1624 77108 sshd: apeace@pts/0 0.6 0.0 1544 9940 redis-server /etc/redis/redis.conf 0.6 0.0 1524 25848 /usr/sbin/ntpd -p /var/run/ntpd.pid -g -u 103:105 0.5 0.0 1324 119880 rsyslogd -c4 0.4 0.0 1084 49308 /usr/sbin/sshd 0.4 0.0 1028 44376 /usr/sbin/exim4 -bd -q30m 0.3 0.0 904 6876 ps -eo pmem,pcpu,rss,vsize,args 0.3 0.0 888 21124 cron 0.3 0.0 868 23472 dbus-daemon --system --fork 0.2 0.0 732 19624 -bash 0.2 0.0 628 6128 /sbin/getty -8 38400 tty1 0.2 0.0 628 16952 upstart-udev-bridge --daemon 0.2 0.0 564 16800 udevd --daemon 0.2 0.0 552 16796 udevd --daemon 0.2 0.0 548 16796 udevd --daemon 0.0 0.0 0 0 [xenwatch] 0.0 0.0 0 0 [xenbus] 0.0 0.0 0 0 [sync_supers] 0.0 0.0 0 0 [netns] 0.0 0.0 0 0 [migration/3] 0.0 0.0 0 0 [migration/2] 0.0 0.0 0 0 [migration/1] 0.0 0.0 0 0 [migration/0] 0.0 0.0 0 0 [kthreadd] 0.0 0.0 0 0 [kswapd0] 0.0 0.0 0 0 [kstriped] 0.0 0.0 0 0 [ksoftirqd/3] 0.0 0.0 0 0 [ksoftirqd/2] 0.0 0.0 0 0 [ksoftirqd/1] 0.0 0.0 0 0 [ksoftirqd/0] 0.0 0.0 0 0 [ksnapd] 0.0 0.0 0 0 [kseriod] 0.0 0.0 0 0 [kjournald] 0.0 0.0 0 0 [khvcd] 0.0 0.0 0 0 [khelper] 0.0 0.0 0 0 [kblockd/3] 0.0 0.0 0 0 [kblockd/2] 0.0 0.0 0 0 [kblockd/1] 0.0 0.0 0 0 [kblockd/0] 0.0 0.0 0 0 [flush-202:1] 0.0 0.0 0 0 [events/3] 0.0 0.0 0 0 [events/2] 0.0 0.0 0 0 [events/1] 0.0 0.0 0 0 [events/0] 0.0 0.0 0 0 [crypto/3] 0.0 0.0 0 0 [crypto/2] 0.0 0.0 0 0 [crypto/1] 0.0 0.0 0 0 [crypto/0] 0.0 0.0 0 0 [cpuset] 0.0 0.0 0 0 [bdi-default] 0.0 0.0 0 0 [async/mgr] 0.0 0.0 0 0 [aio/3] 0.0 0.0 0 0 [aio/2] 0.0 0.0 0 0 [aio/1] 0.0 0.0 0 0 [aio/0] Now, I know that ps is not the best for viewing process memory usage, but that's because it tends to report more memory than is actually being used...meaning no matter how you look at it all my processes combined shouldn't be using near 122mb, even if you account for the disk cache. What's more, memory usage is growing all the time. I've had to restart my server once a week, because once my 256mb fills up it starts swapping, which it wouldn't do just for disk cache. Shouldn't there be some way for me to see the culprit?! I'm new to server admin, so please if there's something obvious I'm overlooking point it out to me. Just for good measure, the output of cat /proc/meminfo: MemTotal: 251140 kB MemFree: 124604 kB Buffers: 20536 kB Cached: 66136 kB SwapCached: 0 kB Active: 65004 kB Inactive: 37576 kB Active(anon): 15932 kB Inactive(anon): 164 kB Active(file): 49072 kB Inactive(file): 37412 kB Unevictable: 0 kB Mlocked: 0 kB SwapTotal: 524284 kB SwapFree: 524284 kB Dirty: 8 kB Writeback: 0 kB AnonPages: 15916 kB Mapped: 10668 kB Shmem: 188 kB Slab: 18604 kB SReclaimable: 10088 kB SUnreclaim: 8516 kB KernelStack: 536 kB PageTables: 1444 kB NFS_Unstable: 0 kB Bounce: 0 kB WritebackTmp: 0 kB CommitLimit: 649852 kB Committed_AS: 64224 kB VmallocTotal: 34359738367 kB VmallocUsed: 752 kB VmallocChunk: 34359737600 kB DirectMap4k: 262144 kB DirectMap2M: 0 kB EDIT: I had misinterpreted the meaning of free -m at first. But even so: the important thing is that my OS eventually begins to use swap memory if I don't restart my server, which disk caching wouldn't do. So where do I look to see what is using all this memory?

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  • Where's the Swap File/Partition?

    - by chrisbunney
    I'm investigating the virtual memory configuration of a Debian based Amazon EC2 instance, and as my background isn't in system admin, I'm slightly confused by what I'm seeing. We're using MongoDB, and the monitoring server we have indicates that the Mongo process is using about 20GB of swap space, however I can't figure out where this is located on the server. As far as I can tell from using the various suggested methods from Google, there is either a much smaller amount, or none at all. top indicates that there is 1.8GB of swap memory: top - 15:35:21 up 6 days, 3:23, 1 user, load average: 1.60, 1.43, 1.37 Tasks: 47 total, 2 running, 45 sleeping, 0 stopped, 0 zombie Cpu(s): 0.0%us, 1.3%sy, 0.0%ni, 14.7%id, 83.8%wa, 0.0%hi, 0.0%si, 0.1%st Mem: 3928924k total, 2855572k used, 1073352k free, 640564k buffers Swap: 0k total, 0k used, 0k free, 1887788k cached swapon -s doesn't seem to think there's any swap space: Filename Type Size Used Priority free -m doesn't think there's any swap either: total used free shared buffers cached Mem: 3836 3663 172 0 626 2701 -/+ buffers/cache: 336 3500 Swap: 0 0 0 And neither does vmstat: procs -----------memory---------- ---swap-- -----io---- -system-- ----cpu---- r b swpd free buff cache si so bi bo in cs us sy id wa 0 3 0 66224 641372 2874744 0 0 21 5012 21 33 2 2 76 19 But cat /etc/fstab thinks there is a swap partition: /dev/xvda1 / ext3 defaults 1 1 /dev/xvda2 /mnt ext3 defaults 0 0 /dev/xvda3 swap swap defaults 0 0 none /proc proc defaults 0 0 none /sys sysfs defaults 0 0 However df -k gives no indication of the xvda3 partition: Filesystem 1K-blocks Used Available Use% Mounted on /dev/xvda1 16513960 15675324 0 100% / tmpfs 1964460 8 1964452 1% /lib/init/rw udev 1914148 28 1914120 1% /dev tmpfs 1964460 4 1964456 1% /dev/shm So I really don't know what to make of this, because I appear to have a process using about 10 times more virtual memory than what might be available, and I have no idea where this virtual memory is on the system. I'm probably misinterpreting the output of the tools, so I'd be grateful if someone would be able to set me straight: What have I got wrong, what's the right interpretation, and how do you reach that interpretation? EDIT0: We use 10gen's MMS for monitoring the database, the relevant section for memory from the last data point is: "mem": { "virtual": 20749, "bits": 64, "supported": true, "mappedWithJournal": 20376, "mapped": 10188, "resident": 1219 }, This JSON is specific to the database process (I believe) rather than the system as a whole. fdisk -l /dev/xvda outputs... nothing? I tried each of the 3 xvda entries in /etc/fstab as well: root@ip:~# fdisk -l /dev/xvda1 Disk /dev/xvda1: 34.4 GB, 34359738368 bytes 255 heads, 63 sectors/track, 4177 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00000000 Disk /dev/xvda1 doesn't contain a valid partition table root@ip:~# fdisk -l /dev/xvda2 root@ip:~# fdisk -l /dev/xvda3 root@ip:~# Edit1: Output of cat /proc/meminfo for the sake of completeness: MemTotal: 3928924 kB MemFree: 726600 kB Buffers: 648368 kB Cached: 2216556 kB SwapCached: 0 kB Active: 1945100 kB Inactive: 994016 kB Active(anon): 60476 kB Inactive(anon): 12952 kB Active(file): 1884624 kB Inactive(file): 981064 kB Unevictable: 0 kB Mlocked: 0 kB SwapTotal: 0 kB SwapFree: 0 kB Dirty: 387180 kB Writeback: 0 kB AnonPages: 73380 kB Mapped: 1188260 kB Shmem: 48 kB Slab: 149768 kB SReclaimable: 146076 kB SUnreclaim: 3692 kB KernelStack: 1104 kB PageTables: 16096 kB NFS_Unstable: 0 kB Bounce: 0 kB WritebackTmp: 0 kB CommitLimit: 1964460 kB Committed_AS: 305572 kB VmallocTotal: 34359738367 kB VmallocUsed: 16760 kB VmallocChunk: 34359721448 kB HardwareCorrupted: 0 kB HugePages_Total: 0 HugePages_Free: 0 HugePages_Rsvd: 0 HugePages_Surp: 0 Hugepagesize: 2048 kB DirectMap4k: 3932160 kB DirectMap2M: 0 kB

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  • Courier Maildrop error user unknown. Command output: Invalid user specified

    - by cad
    Hello I have a problem with maildrop. I have read dozens of webs/howto/emails but couldnt solve it. My objective is moving automatically spam messages to a spam folder. My email server is working perfectly. It marks spam in subject and headers using spamassasin. My box has: Ubuntu 9.04 Web: Apache2 + Php5 + MySQL MTA: Postfix 2.5.5 + SpamAssasin + virtual users using mysql IMAP: Courier 0.61.2 + Courier AuthLib WebMail: SquirrelMail I have read that I could use Squirrelmail directly (not a good idea), procmail or maildrop. As I already have maildrop in the box (from courier) I have configured the server to use maildrop (added an entry in transport table for a virtual domain). I found this error in email: This is the mail system at host foo.net I'm sorry to have to inform you that your message could not be delivered to one or more recipients. It's attached below. For further assistance, please send mail to postmaster. If you do so, please include this problem report. You can delete your own text from the attached returned message. The mail system <[email protected]>: user unknown. Command output: Invalid user specified. Final-Recipient: rfc822; [email protected] Action: failed Status: 5.1.1 Diagnostic-Code: x-unix; Invalid user specified. ---------- Forwarded message ---------- From: test <[email protected]> To: [email protected] Date: Sat, 1 May 2010 19:49:57 +0100 Subject: fail fail An this in the logs May 1 18:50:18 foo.net postfix/smtpd[14638]: connect from mail-bw0-f212.google.com[209.85.218.212] May 1 18:50:19 foo.net postfix/smtpd[14638]: 8A9E9DC23F: client=mail-bw0-f212.google.com[209.85.218.212] May 1 18:50:19 foo.net postfix/cleanup[14643]: 8A9E9DC23F: message-id=<[email protected]> May 1 18:50:19 foo.net postfix/qmgr[14628]: 8A9E9DC23F: from=<[email protected]>, size=1858, nrcpt=1 (queue active) May 1 18:50:23 foo.net postfix/pickup[14627]: 1D4B4DC2AA: uid=5002 from=<[email protected]> May 1 18:50:23 foo.net postfix/cleanup[14643]: 1D4B4DC2AA: message-id=<[email protected]> May 1 18:50:23 foo.net postfix/pipe[14644]: 8A9E9DC23F: to=<[email protected]>, relay=spamassassin, delay=3.8, delays=0.55/0.02/0/3.2, dsn=2.0.0, status=sent (delivered via spamassassin service) May 1 18:50:23 foo.net postfix/qmgr[14628]: 8A9E9DC23F: removed May 1 18:50:23 foo.net postfix/qmgr[14628]: 1D4B4DC2AA: from=<[email protected]>, size=2173, nrcpt=1 (queue active) **May 1 18:50:23 foo.netpostfix/pipe[14648]: 1D4B4DC2AA: to=<[email protected]>, relay=maildrop, delay=0.22, delays=0.06/0.01/0/0.15, dsn=5.1.1, status=bounced (user unknown. Command output: Invalid user specified. )** May 1 18:50:23 foo.net postfix/cleanup[14643]: 4C2BFDC240: message-id=<[email protected]> May 1 18:50:23 foo.net postfix/qmgr[14628]: 4C2BFDC240: from=<>, size=3822, nrcpt=1 (queue active) May 1 18:50:23 foo.net postfix/bounce[14651]: 1D4B4DC2AA: sender non-delivery notification: 4C2BFDC240 May 1 18:50:23 foo.net postfix/qmgr[14628]: 1D4B4DC2AA: removed May 1 18:50:24 foo.net postfix/smtp[14653]: 4C2BFDC240: to=<[email protected]>, relay=gmail-smtp-in.l.google.com[209.85.211.97]:25, delay=0.91, delays=0.02/0.03/0.12/0.74, dsn=2.0.0, status=sent (250 2.0.0 OK 1272739824 37si5422420ywh.59) May 1 18:50:24 foo.net postfix/qmgr[14628]: 4C2BFDC240: removed My config files: http://lar3d.net/main.cf (/etc/postfix) http://lar3d.net/master.c (/etc/postfix) http://lar3d.net/local.cf (/etc/spamassasin) http://lar3d.net/maildroprc (maildroprc) If I change master.cf line (as suggested here) maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d ${recipient} with maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d vmail ${recipient} I get the email in /home/vmail/MailDir instead of the correct dir (/home/vmail/foo.net/info/.SPAM ) After reading a lot I have some guess but not sure. - Maybe I have to install userdb? - Maybe is something related with mysql, but everything is working ok - If I try with procmail I will face same problem... - What are flags DRhu for? Couldnt find doc about them - In some places I found maildrop line with more parameters flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d $ ${recipient} ${extension} ${recipient} ${user} ${nexthop} ${sender} I am really lost. Dont know how to continue. If you have any idea or need another config file please let me know. Thanks!!!

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  • What do these messages in the Qmail maillog indicate?

    - by Griffo
    There seems to be an endless supply of messages in the Qmail maillog for a single address. Can anyone shed some light on why this might be and whether it is a problem? To me it looks like either spam or some sort of unhandled problem. It strikes me as unusual that the 'from=' field is blank. This is on a VPS using Plesk in case that's important. Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23593]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23586]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23586]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23585]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23585]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23584]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23584]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23583]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23583]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23600]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23600]: [email protected] Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23599]: from= Jun 30 15:10:17 vps-1001108-595 qmail-remote-handlers[23599]: [email protected] EDIT Here's a sample of one of the emails: Received: (qmail 5603 invoked for bounce); 29 Jun 2011 07:46:31 +0100 Date: 29 Jun 2011 07:46:31 +0100 From: [email protected] To: [email protected] Subject: failure notice Hi. This is the qmail-send program at vps-1001108-595.cp.blacknight.com. I'm afraid I wasn't able to deliver your message to the following addresses. This is a permanent error; I've given up. Sorry it didn't work out. <[email protected]>: 200.147.36.13 does not like recipient. Remote host said: 450 4.7.1 Client host rejected: cannot find your hostname, [78.153.208.195] Giving up on 200.147.36.13. I'm not going to try again; this message has been in the queue too long. --- Below this line is a copy of the message. Return-Path: <[email protected]> Received: (qmail 15585 invoked by uid 48); 22 Jun 2011 07:38:26 +0100 Date: 22 Jun 2011 07:38:26 +0100 Message-ID: <[email protected]> To: [email protected] Subject: Cadastre-se e Concorra ? um Carro! MIME-Version: 1.0 Content-type: text/html; charset=iso-8859-1 From: Cielo Fidelidade <[email protected]> <!DOCTYPE HTML> <html> ... <body text removed> <html> If I understand this correctly, this is saying that an email sent by my server, from address [email protected], could not be delivered. However, [email protected] is not a valid email address on my server, so how can email be sent from this address on my server? I have tested whether my server is acting as an open relay, and it isn't. So how else could this be happening? I am getting thousands of these every day. What can I do to prevent it?

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  • Postfix MySql Dovecot - SMTP Authentication Failure

    - by borncamp
    Hello I have a Postfix setup with Dovecot and MySql. The server is running Debian Squeeze. The MySql server is a slave that has data pushed to it from a primary (postfix) mail server(running a different os). The emails are stored on a replicated GlusterFS volume. I am able to check email using thunderbird over IMAP. However, SMTP requests fail. After turning on query logs for the MySql server I have noticed that no query statement is executed to retrieve the user information when an SMTP client tries to authenticate. I'd like to know what I'm doing wrong or what the next troubleshooting steps are. I'm about to pull my hair out. Below is some log and configuration data that I thought would be relevant. You're help is much obliged. The file /var/log/mail.log shows Oct 11 14:54:16 mailbox2 postfix/smtpd[25017]: connect from unknown[192.168.0.44] Oct 11 14:54:19 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL PLAIN authentication failed: Oct 11 14:54:25 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL LOGIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:48 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL PLAIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:54 mailbox2 postfix/smtpd[25017]: warning: unknown[192.168.0.44]: SASL LOGIN authentication failed: VXNlcm5hbWU6 Oct 11 14:55:57 mailbox2 postfix/smtpd[25017]: disconnect from unknown[192.168.0.44] This is my dovecot.conf file log_timestamp = "%Y-%m-%d %H:%M:%S " mail_location = maildir:/var/mail/virtual/%d/%n/ auth_mechanisms = plain login disable_plaintext_auth = no namespace { inbox = yes location = prefix = INBOX. separator = . type = private } passdb { args = /etc/dovecot/dovecot-mysql.conf driver = sql } protocols = imap pop3 service auth { unix_listener /var/spool/postfix/private/auth { group = postfix mode = 0660 user = postfix } unix_listener auth-master { mode = 0600 user = postfix } user = root } ssl_cert = </etc/ssl/certs/dovecot.pem ssl_key = </etc/ssl/private/dovecot.pem userdb { args = /etc/dovecot/dovecot-mysql.conf driver = sql } protocol lda { auth_socket_path = /var/run/dovecot/auth-master mail_plugins = sieve postmaster_address = [email protected] } protocol pop3 { pop3_uidl_format = %08Xu%08Xv } Here is my dovecot-mysql.conf file: connect = host=127.0.0.1 dbname=postfix user=postfix password=ffjM2MYAqQtAzRHX driver = mysql default_pass_scheme = MD5-CRYPT password_query = SELECT username AS user,password FROM mailbox WHERE username = '%u' AND active='1' user_query = SELECT CONCAT('/var/mail/virtual/', maildir) AS home, 1001 AS uid, 109 AS gid, CONCAT('*:messages=10000:bytes=',quota) as quota_rule, 'Trash:ignore' AS quota_rule2 FROM mailbox WHERE username = '%u' AND active='1' Here is my output from 'postconf -n': append_dot_mydomain = no biff = no bounce_template_file = /etc/postfix/bounce.cf broken_sasl_auth_clients = yes config_directory = /etc/postfix delay_warning_time = 0h dovecot_destination_recipient_limit = 1 inet_interfaces = all local_recipient_maps = $virtual_mailbox_maps local_transport = virtual mailbox_command = procmail -a "$EXTENSION" mailbox_size_limit = 0 maximal_queue_lifetime = 1d message_size_limit = 25600000 mydestination = mailbox2.cws.net, debian.local.cws.net, localhost.local.cws.net, localhost myhostname = mailbox2.cws.net mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 172.18.0.119 63.164.138.3 myorigin = /etc/mailname proxy_read_maps = $local_recipient_maps $mydestination $virtual_alias_maps $virtual_alias_domains $virtual_mailbox_maps $virtual_mailbox_domains $relay_recipient_maps $relay_domains $canonical_maps $sender_canonical_maps $recipient_canonical_maps $relocated_maps $transport_maps $mynetworks $virtual_mailbox_limit_maps readme_directory = no recipient_delimiter = + relay_domains = relayhost = smtp_connect_timeout = 10 smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache smtpd_banner = $myhostname ESMTP $mail_name (Debian/GNU) smtpd_client_message_rate_limit = 50 smtpd_client_recipient_rate_limit = 500 smtpd_client_restrictions = permit_sasl_authenticated, permit_mynetworks smtpd_delay_reject = yes smtpd_discard_ehlo_keyword_address_maps = hash:/etc/postfix/discard_ehlo smtpd_helo_required = yes smtpd_helo_restrictions = permit_mynetworks, reject_invalid_helo_hostname, permit smtpd_recipient_restrictions = permit_mynetworks,permit_sasl_authenticated,reject_unauth_destination smtpd_sasl_auth_enable = yes smtpd_sasl_authenticated_header = yes smtpd_sasl_path = private/auth smtpd_sasl_security_options = noanonymous smtpd_sasl_tls_security_options = $smtpd_sasl_security_options smtpd_sasl_type = dovecot smtpd_sender_restrictions = permit_mynetworks, reject_non_fqdn_sender, reject_unknown_sender_domain, permit smtpd_tls_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_key_file = /etc/ssl/private/ssl-cert-snakeoil.key smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtpd_use_tls = yes transport_maps = hash:/etc/postfix/transport virtual_alias_maps = proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_catchall_maps.cf virtual_gid_maps = static:1001 virtual_mailbox_base = /var/mail/virtual/ virtual_mailbox_domains = proxy:mysql:/etc/postfix/sql/mysql_virtual_domains_maps.cf virtual_mailbox_maps = proxy:mysql:/etc/postfix/sql/mysql_virtual_mailbox_maps.cf, proxy:mysql:/etc/postfix/sql/mysql_virtual_alias_domain_mailbox_maps.cf virtual_transport = dovecot virtual_uid_maps = static:1001

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  • Capistrano + Nginx + Passenger = 403

    - by slimchrisp
    I asked this over at stackoverflow as well, but still haven't received any answers that have helped me to solve this problem. I have spent almost a week at this point trying to solve the issue, and I'm just not making any headway. It seems that this issue is pretty common, but none of the solutions I found online work for me. A buddy of mine is actually creating the same setup, and he is having the same issue. After a few days stuck with the 403 error I started over using this tutorial: http://blog.ninjahideout.com/posts/a-guide-to-a-nginx-passenger-and-rvm-server I had hoped starting from scratch using this tutorial would work, but no dice. Either way, if you view the tutorial you can see what steps I have taken. Here is essentially what I have going on. I have a VPS account on linode.com Server OS is Ubuntu 10.04 Local OS (shouldn't matter, but just so you know) used to deploy with Capistrano is Snow Leopard 10.6.6 I use RVM on the server. Version is 1.2.2 I was previously on ruby-1.9.2-p0 [ i386 ], but per the tutorial listed above I switched to ree-1.8.7-2010.02 [ i386 ]. Running 'which ruby' from the command line verifies that I am using 1.8.7 with the following output: /usr/local/rvm/rubies/ree-1.8.7-2010.02/bin/ruby passenger -v prints the following: Phusion Passenger version 3.0.2 Running 'nginx -v' gives me a message that the command nginx could not be found. The server is definitely there and running as I can use nginx to serve static files, but this could have something to do with my problem. I have two users dealing with the install. root which I used to install everything, and deployer which is a user I created specifically to for deploying my applications My web app directory is in the deployer user's home directory as follows: /home/deployer/webapps/mysite.com/public Per Capistrano default deploy, a symbolic link called current is created in the public folder, and points to /home/deployer/webapps/mysite.com/public/releases/most_current_release I have chmodded the deployer directory recursively to 777 /opt/nginx permissions: rwxr-xr-x /usr/local/rvm/gems/ree-1.8.7-2010.02/gems/passenger-3.0.2 permissions: rwxrwsrwx My nginx config file has gone through just short of eternity variations, but currently looks like this: ================================================================================== worker_processes 1; events { worker_connections 1024; } http { passenger_root /usr/local/rvm/gems/ree-1.8.7-2010.02/gems/passenger-3.0.2; passenger_ruby /usr/local/rvm/bin/passenger_ruby; include mime.types; default_type application/octet-stream; sendfile on; keepalive_timeout 65; server { # listen *:80; server_name mysite.com www.mysite.com; root /home/deployer/webapps/mysite.com/public/current; passenger_enabled on; passenger_friendly_error_pages on; access_log logs/mysite.com/server.log; error_log logs/mysite.com/error.log info; error_page 500 502 503 504 /50x.html; location = /50x.html { root html; } } } ================================================================================== I bounce nginx, hit the site, and boom. 403, and logs say directory index of /home/deployer... is forbidden As others with a similar problem have said, you can drop an index.html into the public/releases/current_release and it will render. But rails no worky. That's basically it. At this point I have just about completely exhausted every possible solution attempt I can think of. I am a programmer and definitely not a sysadmin, so I am 99% sure this has something to do with permissions that I have hosed, but for the life of me I just can't figure out where. If anyone can help I would really really appreciate it. If there's any specific permission things you want me to check (ie groups/permissions), can you please include the commands to do so as well. Hopefully this will help others in the future who read this post. Let me know if there is any other information I can provide, and thanks in advance!!!

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  • 1600+ 'postfix-queue' processes - OK to have this many?

    - by atomicguava
    I have a Plesk 9.5.4 CentOS server running Postfix. I had been having massive problems with the mailq being full of 'double-bounce' email messages containing errors relating to 'Queue File Write Error', but I believe these are now fixed thanks to this thread. My new problem is that when I run top, I can see lots of processes called 'postfix-queue' and have fairly high load: top - 13:59:44 up 6 days, 21:14, 1 user, load average: 2.33, 2.19, 1.96 Tasks: 1743 total, 1 running, 1742 sleeping, 0 stopped, 0 zombie Cpu(s): 5.1%us, 8.8%sy, 0.0%ni, 85.3%id, 0.8%wa, 0.0%hi, 0.0%si, 0.0%st Mem: 3145728k total, 1950640k used, 1195088k free, 0k buffers Swap: 0k total, 0k used, 0k free, 0k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 1324 apache 16 0 344m 33m 5664 S 21.7 1.1 0:03.17 httpd 32443 apache 15 0 350m 36m 6864 S 14.4 1.2 0:13.83 httpd 1678 root 15 0 13948 2568 952 R 2.0 0.1 0:00.37 top 1890 mysql 15 0 689m 318m 7600 S 1.0 10.4 219:45.23 mysqld 1394 apache 15 0 352m 41m 5972 S 0.7 1.3 0:03.91 httpd 1369 apache 15 0 344m 33m 5444 S 0.3 1.1 0:02.03 httpd 1592 apache 15 0 349m 37m 5912 S 0.3 1.2 0:02.52 httpd 1633 apache 15 0 336m 20m 1828 S 0.3 0.7 0:00.01 httpd 1952 root 19 0 335m 28m 10m S 0.3 0.9 1:35.41 httpd 1 root 15 0 10304 732 612 S 0.0 0.0 0:04.41 init 1034 mhandler 15 0 11520 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1036 mhandler 15 0 11516 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1041 mhandler 17 0 11516 1156 884 S 0.0 0.0 0:00.00 postfix-queue 1043 mhandler 15 0 11512 1116 860 S 0.0 0.0 0:00.00 postfix-queue 1063 mhandler 16 0 11516 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1068 mhandler 15 0 11516 1128 860 S 0.0 0.0 0:00.00 postfix-queue 1071 mhandler 17 0 11512 1152 884 S 0.0 0.0 0:00.00 postfix-queue 1072 mhandler 15 0 11512 1116 860 S 0.0 0.0 0:00.00 postfix-queue 1081 mhandler 16 0 11516 1156 884 S 0.0 0.0 0:00.00 postfix-queue 1082 mhandler 15 0 11512 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1089 popuser 15 0 33892 1972 1200 S 0.0 0.1 0:00.02 pop3d 1116 mhandler 16 0 11516 1164 884 S 0.0 0.0 0:00.00 postfix-queue 1117 mhandler 15 0 11516 1124 860 S 0.0 0.0 0:00.00 postfix-queue 1120 mhandler 16 0 11516 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1121 mhandler 15 0 11512 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1130 mhandler 17 0 11516 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1131 mhandler 15 0 11516 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1149 root 17 -4 12572 680 356 S 0.0 0.0 0:00.00 udevd 1181 mhandler 16 0 11516 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1183 mhandler 15 0 11512 1116 860 S 0.0 0.0 0:00.00 postfix-queue 1224 mhandler 16 0 11516 1160 884 S 0.0 0.0 0:00.00 postfix-queue 1225 mhandler 15 0 11516 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1228 apache 15 0 345m 34m 5472 S 0.0 1.1 0:04.64 httpd 1241 mhandler 16 0 11516 1156 884 S 0.0 0.0 0:00.00 postfix-queue 1242 mhandler 15 0 11512 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1251 mhandler 17 0 11516 1156 884 S 0.0 0.0 0:00.00 postfix-queue 1252 mhandler 15 0 11516 1120 860 S 0.0 0.0 0:00.00 postfix-queue 1258 apache 15 0 349m 37m 5444 S 0.0 1.2 0:01.28 httpd When I run ps -Al | grep -c postfix-queue it returns 1618! My question is this: is this normal or is there something else going wrong with Postfix? Right now, if I run mailq it is empty, and qshape deferred / qshape active are empty too. Thanks in advance for your help.

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  • Network update solutions for a company of ~20 (5 local, 15 remote)?

    - by Margaret
    Hi all This is probably going to be a bit up in the air, because we're still in the "reaching towards solutions" phase, but I figured I'd see what you guys had to say. Plus I honestly know very little about systems and what is good and bad pratice. My organisation has always more or less worked on the concept of local machines; since it primarily employed contractors who were working from home, each of those people was largely responsible for their own machine and backup procedures and the like. We're now expanding, though we're still reasonably small (we're up to about 20 staff members). Most people still work remotely, but we have a central office where about five people are working. But we're getting large enough that we're starting to think it would be a good idea to have a central file server, and things like that - if someone gets hit by a bus, we want someone else to know where to look for the files to continue their work. A lot of the people who work for us remotely work on projects for other companies as well, so I don't want to force them to log in to our server whenever they're on a network. But I do want to make connection to be as painless as possible to do so, to improve utilisation. The other thing is that we're getting more people who would like to remote into the office server and do their work there. Our current remote connection application is an SSH install that allows people access to the network; the problem is, it's a black box to me, and I've never understood how to even connect to it (despite supposedly being de facto sysadmin). Thus far I've been able to bounce questions about how to get it working to the guy who does know it well, but he's leaving the company soon. So we probably need a solution for this that I actually understand. We were knocking around the idea of implementing a VPN with some form of remote desktop, and someone mentioned that this was largely a matter of purchasing a router capable of it; I'm not sure of the truth of that statement. This is what we have in the office: Two shiny new i7 servers, each running Windows Server 2008. Precise eventual layout is still being debated, a little, but the current suggestion is that one is primary database crunching, while the other is a warm backup of the databases, along with running Reporting Services. They currently have SQL Server 2008 installed on them, which is being connected to via the 'sa' account. We're hoping to make each person use their own account (preferably one tied to the 'central' password we set up, so we can use Windows Authentication). An older server, running XP Pro, that we are currently using as a test bed for a project that requires access to older versions of software. This machine is also being used to take backups, but I'm thinking of moving that functionality elsewhere. A spare desktop from a guy who left the company (XP Pro). We're thinking of bumping up the hard disk space and using it as the magical file server that's going to solve one particular everything. Assorted desktops, laptops, etc, at least one for each person in the office (mix of Win XP and Win 7; occasionally a person who normally works remotely might drop in to the office and bring a laptop bearing Vista, but it's pretty rare). All are set up as local user accounts at the moment; I don't know if it's the best arrangement. Purchasing more hardware is not a big problem, but we figure we might as well make use of what we've got first. Is Active Directory a big magic wand that's going to solve all the world's problems? Is there some other arrangement we should be looking to instead?

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  • Changing CSS with jQuery syntax in Silverlight using jLight

    - by Timmy Kokke
    Lately I’ve ran into situations where I had to change elements or had to request a value in the DOM from Silverlight. jLight, which was introduced in an earlier article, can help with that. jQuery offers great ways to change CSS during runtime. Silverlight can access the DOM, but it isn’t as easy as jQuery. All examples shown in this article can be looked at in this online demo. The code can be downloaded here.   Part 1: The easy stuff Selecting and changing properties is pretty straight forward. Setting the text color in all <B> </B> elements can be done using the following code:   jQuery.Select("b").Css("color", "red");   The Css() method is an extension method on jQueryObject which is return by the jQuery.Select() method. The Css() method takes to parameters. The first is the Css style property. All properties used in Css can be entered in this string. The second parameter is the value you want to give the property. In this case the property is “color” and it is changed to “red”. To specify which element you want to select you can add a :selector parameter to the Select() method as shown in the next example.   jQuery.Select("b:first").Css("font-family", "sans-serif");   The “:first” pseudo-class selector selects only the first element. This example changes the “font-family” property of the first <B></B> element to “sans-serif”. To make use of intellisense in Visual Studio I’ve added a extension methods to help with the pseudo-classes. In the example below the “font-weight” of every “Even” <LI></LI> is set to “bold”.   jQuery.Select("li".Even()).Css("font-weight", "bold");   Because the Css() extension method returns a jQueryObject it is possible to chain calls to Css(). The following example show setting the “color”, “background-color” and the “font-size” of all headers in one go.   jQuery.Select(":header").Css("color", "#12FF70") .Css("background-color", "yellow") .Css("font-size", "25px");   Part 2: More complex stuff In only a few cases you need to change only one style property. More often you want to change an entire set op style properties all in one go.  You could chain a lot of Css() methods together. A better way is to add a class to a stylesheet and define all properties in there. With the AddClass() method you can set a style class to a set of elements. This example shows how to add the “demostyle” class to all <B></B> in the document.   jQuery.Select("b").AddClass("demostyle");   Removing the class works in the same way:   jQuery.Select("b").RemoveClass("demostyle");   jLight is build for interacting with to the DOM from Silverlight using jQuery. A jQueryObjectCss object can be used to define different sets of style properties in Silverlight. The over 60 most common Css style properties are defined in the jQueryObjectCss class. A string indexer can be used to access all style properties ( CssObject1[“background-color”] equals CssObject1.BackgroundColor). In the code below, two jQueryObjectCss objects are defined and instantiated.   private jQueryObjectCss CssObject1; private jQueryObjectCss CssObject2;   public Demo2() { CssObject1 = new jQueryObjectCss { BackgroundColor = "Lime", Color="Black", FontSize = "12pt", FontFamily = "sans-serif", FontWeight = "bold", MarginLeft = 150, LineHeight = "28px", Border = "Solid 1px #880000" }; CssObject2 = new jQueryObjectCss { FontStyle = "Italic", FontSize = "48", Color = "#225522" }; InitializeComponent(); }   Now instead of chaining to set all different properties you can just pass one of the jQueryObjectCss objects to the Css() method. In this case all <LI></LI> elements are set to match this object.   jQuery.Select("li").Css(CssObject1); When using the jQueryObjectCss objects chaining is still possible. In the following example all headers are given a blue backgroundcolor and the last is set to match CssObject2.   jQuery.Select(":header").Css(new jQueryObjectCss{BackgroundColor = "Blue"}) .Eq(-1).Css(CssObject2);   Part 3: The fun stuff Having Silverlight call JavaScript and than having JavaScript to call Silverlight requires a lot of plumbing code. Everything has to be registered and strings are passed back and forth to execute the JavaScript. jLight makes this kind of stuff so easy, it becomes fun to use. In a lot of situations jQuery can call a function to decide what to do, setting a style class based on complex expressions for example. jLight can do the same, but the callback methods are defined in Silverlight. This example calls the function() method for each <LI></LI> element. The callback method has to take a jQueryObject, an integer and a string as parameters. In this case jLight differs a bit from the actual jQuery implementation. jQuery uses only the index and the className parameters. A jQueryObject is added to make it simpler to access the attributes and properties of the element. If the text of the listitem starts with a ‘D’ or an ‘M’ the class is set. Otherwise null is returned and nothing happens.   private void button1_Click(object sender, RoutedEventArgs e) { jQuery.Select("li").AddClass(function); }   private string function(jQueryObject obj, int index, string className) { if (obj.Text[0] == 'D' || obj.Text[0] == 'M') return "demostyle"; return null; }   The last thing I would like to demonstrate uses even more Silverlight and less jLight, but demonstrates the power of the combination. Animating a style property using a Storyboard with easing functions. First a dependency property is defined. In this case it is a double named Intensity. By handling the changed event the color is set using jQuery.   public double Intensity { get { return (double)GetValue(IntensityProperty); } set { SetValue(IntensityProperty, value); } }   public static readonly DependencyProperty IntensityProperty = DependencyProperty.Register("Intensity", typeof(double), typeof(Demo3), new PropertyMetadata(0.0, IntensityChanged));   private static void IntensityChanged(DependencyObject d, DependencyPropertyChangedEventArgs e) { var i = (byte)(double)e.NewValue; jQuery.Select("span").Css("color", string.Format("#{0:X2}{0:X2}{0:X2}", i)); }   An animation has to be created. This code defines a Storyboard with one keyframe that uses a bounce ease as an easing function. The animation is set to target the Intensity dependency property defined earlier.   private Storyboard CreateAnimation(double value) { Storyboard storyboard = new Storyboard(); var da = new DoubleAnimationUsingKeyFrames(); var d = new EasingDoubleKeyFrame { EasingFunction = new BounceEase(), KeyTime = KeyTime.FromTimeSpan(TimeSpan.FromSeconds(1.0)), Value = value }; da.KeyFrames.Add(d); Storyboard.SetTarget(da, this); Storyboard.SetTargetProperty(da, new PropertyPath(Demo3.IntensityProperty)); storyboard.Children.Add(da); return storyboard; }   Initially the Intensity is set to 128 which results in a gray color. When one of the buttons is pressed, a new animation is created an played. One to animate to black, and one to animate to white.   public Demo3() { InitializeComponent(); Intensity = 128; }   private void button2_Click(object sender, RoutedEventArgs e) { CreateAnimation(255).Begin(); }   private void button3_Click(object sender, RoutedEventArgs e) { CreateAnimation(0).Begin(); }   Conclusion As you can see jLight can make the life of a Silverlight developer a lot easier when accessing the DOM. Almost all jQuery functions that are defined in jLight use the same constructions as described above. I’ve tried to stay as close as possible to the real jQuery. Having JavaScript perform callbacks to Silverlight using jLight will be described in more detail in a future tutorial about AJAX or eventing.

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  • Bug Triage

    In this blog post brain dump, I'll attempt to describe the process my team tries to follow when dealing with new bug reports (specifically, code defect reports). This is not official Microsoft policy, just the way we do things… if you do things differently and want to share, you can do so at the bottom in the comments (or on your blog).Feature Triage TeamA subset of the feature crew, the triage team (which has representations from the PM, Dev and QA disciplines), looks at all unassigned bugs at regular intervals. This can be weekly or daily (or other frequency) dependent on which part of the product cycle we are in and what the untriaged bug load looks like. They discuss each bug considering the evidence and make a decision of whether the bug goes from Not Yet Assigned to Assigned (plus the name of the DEV to fix this) or whether it goes from Active to Resolved (which means it gets assigned back to the requestor for closure or further debate if they were not present at the triage meeting). Close to critical milestones, the feature triage team needs to further justify bugs they take to additional higher-level triage teams.Bug Opened = Not Yet AssignedSomeone (typically an SDET from the QA team) creates the bug item (e.g. in TFS), ensuring they populate all the relevant fields including: Title, Description, Repro Steps (including the Actual Result at the end of the steps), attachments of code and/or screenshots, Build number that they observed the issue in, regression details if applicable, how it was found, if a test case exists or needs to be created etc. They also indicate their opinion on the Priority and Severity. The bug status is left as Not Yet Assigned."Issue" versus "Fix for issue"The solution to some bugs is easy to determine, e.g. "bug: the column name is misspelled". Obviously the fix is to correct the spelling – still, the triage team should be explicit and enter the correct spelling in the bug's Description. Note that a bad bug name here would be "bug: fix the spelling of the column" (it describes the solution, rather than the problem).Other solutions are trickier to establish, e.g. "bug: the column header is not accessible (can only be clicked on with the mouse, not reached via keyboard)". What is the correct solution here? The last thing to do is leave this undetermined and just assign it to a developer. The solution has to be entered in the description. Behind this type of a bug usually hides a spec defect or a new feature request.The person opening the bug should focus on describing the issue, rather than the solution. The person indicates what the fix is in their opinion by stating the Expected Result (immediately after stating the Actual Result). If they have a complex suggested solution, that should be split out in a separate part, but the triage team has the final say before assigning it. If the solution is lengthy/complicated to describe, the bug can be assigned to the PM. Note: the strict interpretation suggests that any bug with no clear, obvious solution is always a hole in the spec and should always go to the PM. This also ensures the spec gets updated.Not Yet Assigned - Not Yet Assigned (on someone else's plate)If the bug is observed in our feature, but the cause is actually another team, we change the Area Path (which is the way we identify teams in TFS) and leave it as Not Yet Assigned. The triage team may add more comments as appropriate including potentially changing the repro steps. In some cases, we may even resolve the bug in our area path and open a new bug in the area path of the other team.Even though there is no action on a dev on the team, the bug still needs to be tracked. One way of doing this is to implement some notification system that informs the team when the tracked bug changed status; another way is to occasionally run a global query (against all area paths) for bugs that have been opened by a member of the team and follow up with the current owners for stale bugs.Not Yet Assigned - ResolvedThis state transition can only be made by the Feature Triage Team.0. Sometimes the bug description is not clear and in that case it gets Resolved as More Information Needed, so the original requestor can provide it.After understanding what the bug item is about, the first decision is to determine whether it needs to go to a dev.1. If it is a known bug, it gets resolved as "Duplicate" and linked to the existing bug.2. If it is "By Design" it gets resolved as such, indicating that the triage team does not think this is a bug.3. If the bug does not repro on latest bits, it is resolved as "No Repro"4. The most painful: If it is decided that we cannot fix it for this release it gets resolved as "Postponed" or "Won't Fix". The former is typically due to resources and time constraints, while the latter is due to deciding that it is not important enough to consume our resources in any release (yes, not all bugs must be fixed!). For both cases, there are other factors that contribute to the decision such as: existence of a reasonable workaround, frequency we expect users to encounter the issue, dependencies on other team to offer a solution, whether it breaks a core scenario, whether it prohibits customer feedback on a major feature, is it a regression from a previous release, impact of the fix on other partner teams (e.g. User Education, User Experience, Localization/Globalization), whether this is the right fix, does the fix impact performance goals, and last but not least, severity of bug (e.g. loss of customer data, security threat, crash, hang). The bar for fixing a bug goes up as the release date approaches. The triage team becomes hardnosed about which bugs to take, while the developers are busy resolving assigned bugs thus everyone drives for Zero Bug Bounce (ZBB). ZBB is when you have 0 active bugs older than 48 hours.Not Yet Assigned - AssignedIf the bug is something we decide to fix in this release and the solution is known, then it is assigned to a DEV. This is either the developer that will do the work, or a Lead that can further assign it to one of his developer team based on a load balancing algorithm of their choosing.Sometimes, the triage team needs the dev to do some investigation work before deciding whether to take the fix; similarly, the checkin for the fix may be gated on code review by the triage team. In these cases, these instructions are provided in the comments section of the bug and when the developer is done they notify the triage team for final decision.Additionally, a Priority and Severity (from 0 to 4) has to be entered, e.g. a P0 means "drop anything you are doing and fix this now" whereas a P4 is something you get to after all P0,1,2,3 bugs are fixed.From a testing perspective, if the bug was found through ad-hoc testing or an external team, the decision is made whether test cases should be added to avoid future regressions. This is communicated to the QA team.Assigned - ResolvedWhen the developer receives the bug (they should be checking daily for new bugs on their plate looking at bugs in order of priority and from older to newer) they can send it back to triage if the information is not clear. Otherwise, they investigate the bug, setting the Sub Status to "Investigating"; if they cannot make progress, they set the Sub Status to "Blocked" and discuss this with triage or whoever else can help them get unblocked. Once they are unblocked, they set the Sub Status to "Working on Solution"; once they are code complete they send a code review request, setting the Sub Status to "Fix Available". After the iterative code review process is over and everyone is happy with the fix, the developer checks it in and changes the state of the bug from Active (and Assigned to them) to Resolved (and Assigned to someone else).The developer needs to ensure that when the status is changed to Resolved that it is assigned to a QA person. For example, maybe the PM opened the bug, but it should be a QA person that will verify the fix - the developer needs to manually change the assignee in that case. Typically the QA person will send an email to the original requestor notifying them that the fix is verified.Resolved - ??In all cases above, note that the final state was Resolved. What happens after that? The final step should be Closed. The bug is closed once the QA person verifying the fix is happy with it. If the person is not happy, then they change the state from Resolved to Active, thus sending it back to the developer. If the developer and QA person cannot reach agreement, then triage can be brought into it. An easy way to do that is change the status back to Not Yet Assigned with appropriate comments so the triage team can re-review.It is important to note that only QA can close a bug. That means that if the opener of the bug was a PM, when the bug gets resolved by the dev it may land on the PM's plate and after a quick review, the PM would re-assign to an SDET, which is the only role that can close bugs. One exception to this is if the person that filed the bug is external: in that case, we leave it Resolved and assigned to them and also send them a notification that they need to verify the fix. Another exception is if specialized developer knowledge is needed for verifying the bug fix (e.g. it was a refactoring suggestion bug typically not observable by the user) in which case it is fine to have a developer verify the fix, and ideally a different developer to the one that opened the bug.Other links on bug triageA quick search reveals that others have talked about this subject, e.g. here, here, here, here and here.Your take?If you have other best practices your team uses to deal with incoming bug reports, feel free to share in the comments below or on your blog. Comments about this post welcome at the original blog.

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