Search Results

Search found 6258 results on 251 pages for 'trouble'.

Page 224/251 | < Previous Page | 220 221 222 223 224 225 226 227 228 229 230 231  | Next Page >

  • jQuery with SharePoint solutions

    - by KunaalKapoor
    For me jQuery is the 'Plan-B' for everything.And most of my projects include the use of jQuery for something or the other, so I decided to write a small note on what works best while using jQuery along with SharePoint.I prefer to use the jQuery JavaScript library, which is far more robust, easier to use, and allows for plugins. Follow the steps below to add jQuery to your master page. For office 365, the prefered location to add jQuery files is the "Site Asserts" library.Deployment Best PracticesThey are only as good as the context it’s being referenced.  In other words, take into account your world before applying it.Script your deployment options.  Folder in SPD. Use the file system.  Make external references.  The JQuery library is on the Microsoft Ajax Content Delivery Network. You may even choose to publish to and from the document library. (pros and cons to this approach)Reference options when referencing the script.ScriptLink will make sure it’s loaded at the top of the page and only loaded once. You need Visual Studio or SPDContent Editor Web Part (CEWP).  Drop it on the page and it’s there.  Easy but dangerousCustom Actions. Great for global deployments of JQuery.  Loads it on every page. It also works in Sandbox installations.Deployment Maintenance Dont’sDon’t add scripts directly to your Master Page. That’s way too much effort because the pages are hard to maintain.Don’t add scripts directly to the CEWP.  Use a content link instead. That will allow for reuse. If you or someone deletes the CEWP you won’t lose code in the web partSecurity.  Any scripts run with the same privileges of the current user.  In other words, you can’t get in trouble.Development Best PracticesDon’t abuse the DOM.  There are better options to load the DOM without hitting it 1,000 times.User other performance boosters.Try other libraries.  Try some custom codeAvoid String conversionMinify your filesUse CAML to reduce number of returns rowsOnly update your JQuery library AFTER RIGOROUS REGRESSION TESTINGCRUD operations can come with some funSP Services wraps SharePoint’s web services for executionThe Bing SDK is pretty easy to use.  You can add it to your page with a script,  put it into a content editor web part and connect it from the address parameters in a list.Steps:1. Go to jquery.com and download the latest jQuery library to your desktop. You want to get the compressed production version, not the development version.2. Open SharePoint Designer (SPD) and connect to the root level of your site's site collection.In SPD, open the "Style Library" folder. Create a folder named "Scripts" inside of the Style Library. Drag the jQuery library JavaScript file from your desktop into the Scripts folder.In the Scripts folder, create a new JavaScript file and name it (e.g. "actions.js").3. If you are using visual studio add a folder for js, you can create a new folder at the root level or if you prefer more cleaner solutions like me, you can use the layouts folder which cleans out on deactivation/uninstall.4. Within the <head> tag of the master page, add a script reference to the jQuery library just above the content place holder named "PlaceHolderAdditonalPageHead" (and above your custom CSS references, if applicable) as follows:<script src="/Style%20Library/Scripts/{jquery library file}.js" type="text/javascript"></script>Immediately after the jQuery library reference add a script reference to your custom scripts file as follows:<script src="/Style%20Library/Scripts/actions.js" type="text/javascript"></script>Inside your script tag, you can test if jQuery is already defined and if not, then add it to the page.<script type='text/javascript'>  if (typeof jQuery == 'undefined')    document.write('<scr'+'ipt type="text/javascript" src="http://code.jquery.com/jquery-1.6.1.min.js"></sc'+'ript>');</script>For the inquisitive few... Read on if you'd like :)Why jQuery on SharePoiny is AwesomeIt’s all about that visual wow factor.  You can get past that, “But it looks like SharePoint”  Take a long list view and put it into JQuery with pagination, etc and you are the hero.  It’s also about new controls you get with JQuery that you couldn’t do before.Why jQuery with SharePoint should be AwfulAlthough it’s fairly easy to get jQuery up and running. Copy/Paste can cause a problem.  If you don’t understand what it’s doing in the Client Object Model and the Document Object Model then it will do things on your site that were completely unexpected. Many blogs will note workarounds they employed on their sites. Why it’s not working: Debugging “sucks”.You need to develop small blocks of functionality, Test it by putting in some alerts  and console.log. Set breakpoints and monitor the DOM via Firebug and some IE development toolsPerformance - It happens all the time. But you should look at the tradeoffs. More time may give you more functionality.Consistency - ”But it works fine on my computer. So test on many browsers.  Take into account client resourcesHarm the Farm -  You need to code wisely and negatively test.  Don’t be the cause of a DoS attack that’s really JQuery asking for a resource over and over and over again.  So code wisely. Do negative testing. Monitor Server Resources.They also did a demo where JQuery did an endless loop to pull data from a list. It’s a poor decision but also an easy mistake.  They spiked their server resources within a couple seconds and had to shut down the call before it brought it down.ConclusionJQuery is now another tool in your tool kit. You don’t have to use it. Use it where it makes sense and where it helps you get your job done.Don’t abuse it, you will pay for it laterIt will add to page bloat so take that into accountIt can slow your performance

    Read the article

  • Use your own domain email and tired of SPAM? SPAMfighter FTW

    - by Dave Campbell
    I wouldn't post this if I hadn't tried it... and I paid for it myself, so don't anybody be thinking I'm reviewing something someone sent me! Long ago and far away I got very tired of local ISPs and 2nd phone lines and took the plunge and got hooked up to cable... yeah I know the 2nd phone line concept may be hard for everyone to understand, but that's how it was in 'the old days'. To avoid having to change email addresses all the time, I decided to buy a domain name, get minimal hosting, and use that for all email into the house. That way if I changed providers, all the email addresses wouldn't have to change. Of course, about a dozen domains later, I have LOTS of pop email addresses and even an exchange address to my client's server... times have changed. What also has changed is the fact that we get SPAM... 'back in the day' when I was a beta tester for the first ISP in Phoenix, someone tried sending an ad to all of us, and what he got in return for his trouble was a bunch of core dumps that locked up his email... if you don't know what a core dump is, ask your grandfather. But in today's world, we're all much more civilized than that, and as with many things, the criminals seem to have much more rights than we do, so we get inundated with email offering all sorts of wild schemes that you'd have to be brain-dead to accept, but yet... if people weren't accepting them, they'd stop sending them. I keep hoping that survival of the smartest would weed out the mental midgets that respond and then the jumk email stop, but that hasn't happened yet anymore than finding high-quality hearing aids at the checkout line of Safeway because of all the dimwits playing music too loud inside their car... but that's another whole topic and I digress. So what's the solution for all the spam? And I mean *all*... on that old personal email address, I am now getting over 150 spam messages a day! Yes I know that's why God invented the delete key, but I took it on as a challenge, and it's a matter of principle... why should I switch email addresses, or convert from [email protected] to something else, or have all my email filtered through some service just because some A-Hole somewhere has a site up trying to phish Ma & Pa Kettle (ask your grandfather about that too) out of their retirement money? Well... I got an email from my cousin the other day while I was writing yet another email rule, and there was a banner on the bottom of his email that said he was protected by SPAMfighter. SPAMfighter huh.... so I took a look at their site, and found yet one more of the supposed tools to help us. But... I read that they're a Microsoft Gold Partner... and that doesn't come lightly... so I took a gamble and here's what I found: I installed it, and had to do a couple things: 1) SPAMfighter stuffed the SPAMfighter folder into my client's exchange address... I deleted it, made a new SPAMfighter folder where I wanted it to go, then in the SPAMfighter Clients settings for Outlook, I told it to put all spam there. 2) It didn't seem to be doing anything. There's a ribbon button that you can select "Block", and I did that, wondering if I was 'training' it, but it wasn't picking up duplicates 3) I sent email to support, and wrote a post on the forum (not to self: reply to that post). By the time the folks from the home office responded, it was the next day, and first up, SPAMfighter knocked down everything that came through when Outlook opend... two thumbs up! I disabled my 'garbage collection' rule from Outlook, and told Outlook not to use the junk folder thinking it was interfering. 4) Day 2 seemed to go about like Day 1... but I hung in there. 5) Day 3 is now a whole new day... I had left Outlook open and hadn't looked at the PC since sometime late yesterday afternoon, and when I looked this morning, *every bit* of spam was in the SPAMfighter folder!! I'm a new paying customer After watching SPAMfighter work this morning, I've purchased a 1-year license, and I now can sit and watch as emails come in and disappear from my inbox into the SPAMfighter folder. No more continual tweaking of the rules. I've got SPAMfighter set to 'Very Hard' filtering... personally I'd rather pull the few real emails out of the SPAMfighter folder than pull spam out of the real folders. Yes this is simply another way of using the delete key, but you know what? ... it feels good :) Here's a screenshot of the stats after just about 48 hours of being onboard: Note that all the ones blocked by me were during Day 1 and 2... I've blocked none today, and everything is blocked. Stay in the 'Light!

    Read the article

  • Stumbling Through: Visual Studio 2010 (Part II)

    I would now like to expand a little on what I stumbled through in part I of my Visual Studio 2010 post and touch on a few other features of VS 2010.  Specifically, I want to generate some code based off of an Entity Framework model and tie it up to an actual data source.  Im not going to take the easy way and tie to a SQL Server data source, though, I will tie it to an XML data file instead.  Why?  Well, why not?  This is purely for learning, there are probably much better ways to get strongly-typed classes around XML but it will force us to go down a path less travelled and maybe learn a few things along the way.  Once we get this XML data and the means to interact with it, I will revisit data binding to this data in a WPF form and see if I cant get reading, adding, deleting, and updating working smoothly with minimal code.  To begin, I will use what was learned in the first part of this blog topic and draw out a data model for the MFL (My Football League) - I dont want the NFL to come down and sue me for using their name in this totally football-related article.  The data model looks as follows, with Teams having Players, and Players having a position and statistics for each season they played: Note that when making the associations between these entities, I was given the option to create the foreign key but I only chose to select this option for the association between Player and Position.  The reason for this is that I am picturing the XML that will contain this data to look somewhat like this: <MFL> <Position/> <Position/> <Position/> <Team>     <Player>         <Statistic/>     </Player> </Team> </MFL> Statistic will be under its associated Player node, and Player will be under its associated Team node no need to have an Id to reference it if we know it will always fall under its parent.  Position, however, is more of a lookup value that will not have any hierarchical relationship to the player.  In fact, the Position data itself may be in a completely different xml file (something Id like to play around with), so in any case, a player will need to reference the position by its Id. So now that we have a simple data model laid out, I would like to generate two things based on it:  A class for each entity with properties corresponding to each entity property An IO class with methods to get data for each entity, either all instances, by Id or by parent. Now my experience with code generation in the past has consisted of writing up little apps that use the code dom directly to regenerate code on demand (or using tools like CodeSmith).  Surely, there has got to be a more fun way to do this given that we are using the Entity Framework which already has built-in code generation for SQL Server support.  Lets start with that built-in stuff to give us a base to work off of.  Right click anywhere in the canvas of our model and select Add Code Generation Item: So just adding that code item seemed to do quite a bit towards what I was intending: It apparently generated a class for each entity, but also a whole ton more.  I mean a TON more.  Way too much complicated code was generated now that code is likely to be a black box anyway so it shouldnt matter, but we need to understand how to make this work the way we want it to work, so lets get ready to do some stumbling through that text template (tt) file. When I open the .tt file that was generated, right off the bat I realize there is going to be trouble there is no color coding, no intellisense no nothing!  That is going to make stumbling through more like groping blindly in the dark while handcuffed and hopping on one foot, which was one of the alternate titles I was considering for this blog.  Thankfully, the community comes to my rescue and I wont have to cast my mind back to the glory days of coding in VI (look it up, kids).  Using the Extension Manager (Available under the Tools menu), I did a quick search for tt editor in the Online Gallery and quickly found the Tangible T4 Editor: Downloading and installing this was a breeze, and after doing so I got some color coding and intellisense while editing the tt files.  If you will be doing any customizing of tt files, I highly recommend installing this extension.  Next, well see if that is enough help for us to tweak that tt file to do the kind of code generation that we wantDid you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

    Read the article

  • Collision Error

    - by Manji
    I am having trouble with collision detection part of the game. I am using touch events to fire the gun as you will see in the video. Note, the android icon is a temporary graphic for the bullets When ever the user touches (represented by clicks in the video)the bullet appears and kills random sprites. As you can see it never touches the sprites it kills or kill the sprites it does touch. My Question is How do I fix it, so that the sprite dies when the bullet hits it? Collision Code snippet: //Handles Collision private void CheckCollisions(){ synchronized(mSurfaceHolder){ for (int i = sprites.size() - 1; i >= 0; i--){ Sprite sprite = sprites.get(i); if(sprite.isCollision(bullet)){ sprites.remove(sprite); mScore++; if(sprites.size() == 0){ mLevel = mLevel +1; currentLevel++; initLevel(); } break; } } } } Sprite Class Code Snippet: //bounding box left<right and top>bottom int left ; int right ; int top ; int bottom ; public boolean isCollision(Beam other) { // TODO Auto-generated method stub if(this.left>other.right || other.left<other.right)return false; if(this.bottom>other.top || other.bottom<other.top)return false; return true; } EDIT 1: Sprite Class: public class Sprite { // direction = 0 up, 1 left, 2 down, 3 right, // animation = 3 back, 1 left, 0 front, 2 right int[] DIRECTION_TO_ANIMATION_MAP = { 3, 1, 0, 2 }; private static final int BMP_ROWS = 4; private static final int BMP_COLUMNS = 3; private static final int MAX_SPEED = 5; private HitmanView gameView; private Bitmap bmp; private int x; private int y; private int xSpeed; private int ySpeed; private int currentFrame = 0; private int width; private int height; //bounding box left<right and top>bottom int left ; int right ; int top ; int bottom ; public Sprite(HitmanView gameView, Bitmap bmp) { this.width = bmp.getWidth() / BMP_COLUMNS; this.height = bmp.getHeight() / BMP_ROWS; this.gameView = gameView; this.bmp = bmp; Random rnd = new Random(); x = rnd.nextInt(gameView.getWidth() - width); y = rnd.nextInt(gameView.getHeight() - height); xSpeed = rnd.nextInt(MAX_SPEED * 2) - MAX_SPEED; ySpeed = rnd.nextInt(MAX_SPEED * 2) - MAX_SPEED; } private void update() { if (x >= gameView.getWidth() - width - xSpeed || x + xSpeed <= 0) { xSpeed = -xSpeed; } x = x + xSpeed; if (y >= gameView.getHeight() - height - ySpeed || y + ySpeed <= 0) { ySpeed = -ySpeed; } y = y + ySpeed; currentFrame = ++currentFrame % BMP_COLUMNS; } public void onDraw(Canvas canvas) { update(); int srcX = currentFrame * width; int srcY = getAnimationRow() * height; Rect src = new Rect(srcX, srcY, srcX + width, srcY + height); Rect dst = new Rect(x, y, x + width, y + height); canvas.drawBitmap(bmp, src, dst, null); } private int getAnimationRow() { double dirDouble = (Math.atan2(xSpeed, ySpeed) / (Math.PI / 2) + 2); int direction = (int) Math.round(dirDouble) % BMP_ROWS; return DIRECTION_TO_ANIMATION_MAP[direction]; } public boolean isCollision(float x2, float y2){ return x2 > x && x2 < x + width && y2 > y && y2 < y + height; } public boolean isCollision(Beam other) { // TODO Auto-generated method stub if(this.left>other.right || other.left<other.right)return false; if(this.bottom>other.top || other.bottom<other.top)return false; return true; } } Bullet Class: public class Bullet { int mX; int mY; private Bitmap mBitmap; //bounding box left<right and top>bottom int left ; int right ; int top ; int bottom ; public Bullet (Bitmap mBitmap){ this.mBitmap = mBitmap; } public void draw(Canvas canvas, int mX, int mY) { this.mX = mX; this.mY = mY; canvas.drawBitmap(mBitmap, mX, mY, null); } }

    Read the article

  • Part 8: How to name EBS Customizations

    - by volker.eckardt(at)oracle.com
    You might wonder why I am discussing this here. The reason is simple: nearly every project has a bit different naming conventions, which makes a the life always a bit complicated (for developers, but also setup responsible, and also for consultants).  Although we always create a document to describe the technical object naming conventions, I have rarely seen a dedicated document  with functional naming conventions. To be precisely, from my stand point, there should always be one global naming definition for an implementation! Let me discuss some related questions: What is the best convention for the customization reference? How to name database objects (tables, packages etc.)? How to name functional objects like Value Sets, Concurrent Programs, etc. How to separate customizations from standard objects best? What is the best convention for the customization reference? The customization reference is the key you use to reference your customization from other lists, from the project plan etc. Usually it is something like XXHU_CONV_22 (HU=customer abbreviation, CONV=Conversion object #22) or XXFA_DEPRN_RPT_02 (FA=Fixed Assets, DEPRN=Short object group, here depreciation, RPT=Report, 02=2nd report in this area) As this is just a reference (not an object name yet), I would prefer the second option. XX=Customization, FA=Main EBS Module linked (you may have sometimes more, but FA is the main) DEPRN_RPT=Short name to specify the customization 02=a unique number Important here is that the HU isn’t used, because XX is enough to mark a custom object, and the 3rd+4th char can be used by the EBS module short name. How to name database objects (tables, packages etc.)? I was leading different developer teams, and I know that one common way is it to take the Customization reference and add more chars behind to classify the object (like _V for view and _T1 for triggers etc.). The only concern I have with this approach is the reusability. If you name your view XXFA_DEPRN_RPT_02_V, no one will by choice reuse this nice view, as it seams to be specific for this CEMLI. My suggestion is rather to name the view XXFA_DEPRN_PERIODS_V and allow herewith reusability for other CEMLIs (although the view will be deployed primarily with CEMLI package XXFA_DEPRN_RPT_02). (check also one of the following Blogs where I will talk about deployment.) How to name Value Sets, Concurrent Programs, etc. For Value Sets I would go with the same convention as for database objects, starting with XX<Module> …. For Concurrent Programs the situation is a bit different. This “object” is seen and used by a lot of users, and they will search for. In many projects it is common to start again with the company short name, or with XX. My proposal would differ. If you have created your own report and you name it “XX: Invoice Report”, the user has to remember that this report does not start with “I”, it starts with X. Would you like typing an X if you are looking for an Invoice report? No, you wouldn’t! So my advise would be to name it:   “Invoice Report (XXAP)”. Still we know it is custom (because of the XXAP), but the end user will type the key “i” to get it (and will see similar reports starting also with “i”). I hope that the general schema behind has now become obvious. How to separate customizations from standard objects best? I would not have this section here if the naming would not play an important role. Unfortunately, we can not always link a custom application to our own object, therefore the naming is really important. In the file system structure we use our $XXyy_TOP, in JAVA_TOP it is perhaps also “xx” in front. But in the database itself? Although there are different concepts in place, still many implementations are using the standard “apps” approach, means custom objects are stored in the apps schema (which should not cause any trouble). Final advise: review the naming conventions regularly, once a month. You may have to add more! And, publish them! To summarize: Technical and functional customized objects should always follow a naming convention. This naming convention should be project wide, and only one place shall be used to maintain (like in a Wiki). If the name is for the end user, rather put a customization identifier at the end; if it is an internal name, start with XX…

    Read the article

  • Namespaces are obsolete

    - by Bertrand Le Roy
    To those of us who have been around for a while, namespaces have been part of the landscape. One could even say that they have been defining the large-scale features of the landscape in question. However, something happened fairly recently that I think makes this venerable structure obsolete. Before I explain this development and why it’s a superior concept to namespaces, let me recapitulate what namespaces are and why they’ve been so good to us over the years… Namespaces are used for a few different things: Scope: a namespace delimits the portion of code where a name (for a class, sub-namespace, etc.) has the specified meaning. Namespaces are usually the highest-level scoping structures in a software package. Collision prevention: name collisions are a universal problem. Some systems, such as jQuery, wave it away, but the problem remains. Namespaces provide a reasonable approach to global uniqueness (and in some implementations such as XML, enforce it). In .NET, there are ways to relocate a namespace to avoid those rare collision cases. Hierarchy: programmers like neat little boxes, and especially boxes within boxes within boxes. For some reason. Regular human beings on the other hand, tend to think linearly, which is why the Windows explorer for example has tried in a few different ways to flatten the file system hierarchy for the user. 1 is clearly useful because we need to protect our code from bleeding effects from the rest of the application (and vice versa). A language with only global constructs may be what some of us started programming on, but it’s not desirable in any way today. 2 may not be always reasonably worth the trouble (jQuery is doing fine with its global plug-in namespace), but we still need it in many cases. One should note however that globally unique names are not the only possible implementation. In fact, they are a rather extreme solution. What we really care about is collision prevention within our application. What happens outside is irrelevant. 3 is, more than anything, an aesthetical choice. A common convention has been to encode the whole pedigree of the code into the namespace. Come to think about it, we never think we need to import “Microsoft.SqlServer.Management.Smo.Agent” and that would be very hard to remember. What we want to do is bring nHibernate into our app. And this is precisely what you’ll do with modern package managers and module loaders. I want to take the specific example of RequireJS, which is commonly used with Node. Here is how you import a module with RequireJS: var http = require("http"); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } This is of course importing a HTTP stack module into the code. There is no noise here. Let’s break this down. Scope (1) is provided by the one scoping mechanism in JavaScript: the closure surrounding the module’s code. Whatever scoping mechanism is provided by the language would be fine here. Collision prevention (2) is very elegantly handled. Whereas relocating is an afterthought, and an exceptional measure with namespaces, it is here on the frontline. You always relocate, using an extremely familiar pattern: variable assignment. We are very much used to managing our local variable names and any possible collision will get solved very easily by picking a different name. Wait a minute, I hear some of you say. This is only taking care of collisions on the client-side, on the left of that assignment. What if I have two libraries with the name “http”? Well, You can better qualify the path to the module, which is what the require parameter really is. As for hierarchical organization, you don’t really want that, do you? RequireJS’ module pattern does elegantly cover the bases that namespaces used to cover, but it also promotes additional good practices. First, it promotes usage of self-contained, single responsibility units of code through the closure-based, stricter scoping mechanism. Namespaces are somewhat more porous, as using/import statements can be used bi-directionally, which leads us to my second point… Sane dependency graphs are easier to achieve and sustain with such a structure. With namespaces, it is easy to construct dependency cycles (that’s bad, mmkay?). With this pattern, the equivalent would be to build mega-components, which are an easier problem to spot than a decay into inter-dependent namespaces, for which you need specialized tools. I really like this pattern very much, and I would like to see more environments implement it. One could argue that dependency injection has some commonalities with this for example. What do you think? This is the half-baked result of some morning shower reflections, and I’d love to read your thoughts about it. What am I missing?

    Read the article

  • Managing text-maps in a 2D array on to be painted on HTML5 Canvas

    - by weka
    So, I'm making a HTML5 RPG just for fun. The map is a <canvas> (512px width, 352px height | 16 tiles across, 11 tiles top to bottom). I want to know if there's a more efficient way to paint the <canvas>. Here's how I have it right now. How tiles are loaded and painted on map The map is being painted by tiles (32x32) using the Image() piece. The image files are loaded through a simple for loop and put into an array called tiles[] to be PAINTED on using drawImage(). First, we load the tiles... and here's how it's being done: // SET UP THE & DRAW THE MAP TILES tiles = []; var loadedImagesCount = 0; for (x = 0; x <= NUM_OF_TILES; x++) { var imageObj = new Image(); // new instance for each image imageObj.src = "js/tiles/t" + x + ".png"; imageObj.onload = function () { console.log("Added tile ... " + loadedImagesCount); loadedImagesCount++; if (loadedImagesCount == NUM_OF_TILES) { // Onces all tiles are loaded ... // We paint the map for (y = 0; y <= 15; y++) { for (x = 0; x <= 10; x++) { theX = x * 32; theY = y * 32; context.drawImage(tiles[5], theY, theX, 32, 32); } } } }; tiles.push(imageObj); } Naturally, when a player starts a game it loads the map they last left off. But for here, it an all-grass map. Right now, the maps use 2D arrays. Here's an example map. [[4, 1, 4, 1, 4, 1, 4, 1, 4, 1, 4, 1, 1, 1, 1, 1], [1, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 1, 1, 1, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 1, 13, 13, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 1, 13, 13, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 1, 13, 13, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 1, 1, 1, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 13, 13, 13, 1, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 13, 13, 11, 11, 11, 13, 13, 13, 13, 13, 13, 13, 1], [13, 13, 13, 1, 1, 1, 1, 1, 1, 1, 13, 13, 13, 13, 13, 1], [1, 1, 1, 13, 13, 13, 13, 13, 13, 13, 13, 13, 13, 1, 1, 1]]; I get different maps using a simple if structure. Once the 2d array above is return, the corresponding number in each array will be painted according to Image() stored inside tile[]. Then drawImage() will occur and paint according to the x and y and times it by 32 to paint on the correct x-y coordinate. How multiple map switching occurs With my game, maps have five things to keep track of: currentID, leftID, rightID, upID, and bottomID. currentID: The current ID of the map you are on. leftID: What ID of currentID to load when you exit on the left of current map. rightID: What ID of currentID to load when you exit on the right of current map. downID: What ID of currentID to load when you exit on the bottom of current map. upID: What ID of currentID to load when you exit on the top of current map. Something to note: If either leftID, rightID, upID, or bottomID are NOT specific, that means they are a 0. That means they cannot leave that side of the map. It is merely an invisible blockade. So, once a person exits a side of the map, depending on where they exited... for example if they exited on the bottom, bottomID will the number of the map to load and thus be painted on the map. Here's a representational .GIF to help you better visualize: As you can see, sooner or later, with many maps I will be dealing with many IDs. And that can possibly get a little confusing and hectic. The obvious pros is that it load 176 tiles at a time, refresh a small 512x352 canvas, and handles one map at time. The con is that the MAP ids, when dealing with many maps, may get confusing at times. My question Is this an efficient way to store maps (given the usage of tiles), or is there a better way to handle maps? I was thinking along the lines of a giant map. The map-size is big and it's all one 2D array. The viewport, however, is still 512x352 pixels. Here's another .gif I made (for this question) to help visualize: Sorry if you cannot understand my English. Please ask anything you have trouble understanding. Hopefully, I made it clear. Thanks.

    Read the article

  • WMemoryProfiler is Released

    - by Alois Kraus
    What is it? WMemoryProfiler is a managed profiling Api to aid integration testing. This free library can get managed heap statistics and memory usage for your own process (remember testing) and other processes as well. The best thing is that it does work from .NET 2.0 up to .NET 4.5 in x86 and x64. To make it more interesting it can attach to any running .NET process. The reason why I do mention this is that commercial profilers do support this functionality only for their professional editions. An normally only since .NET 4.0 since the profiling API only since then does support attaching to a running process. This thing does differ in many aspects from “normal” profilers because while profiling yourself you can get all objects from all managed heaps back as an object array. If you ever wanted to change the state of an object which does only exist a method local in another thread you can get your hands on it now … Enough theory. Show me some code /// <summary> /// Show feature to not only get statisics out of a process but also the newly allocated /// instances since the last call to MarkCurrentObjects. /// GetNewObjects does return the newly allocated objects as object array /// </summary> static void InstanceTracking() { using (var dumper = new MemoryDumper()) // if you have problems use to see the debugger windows true,true)) { dumper.MarkCurrentObjects(); Allocate(); ILookup<Type, object> newObjects = dumper.GetNewObjects() .ToLookup( x => x.GetType() ); Console.WriteLine("New Strings:"); foreach (var newStr in newObjects[typeof(string)] ) { Console.WriteLine("Str: {0}", newStr); } } } … New Strings: Str: qqd Str: String data: Str: String data: 0 Str: String data: 1 … This is really hot stuff. Not only you can get heap statistics but you can directly examine the new objects and make queries upon them. When I do find more time I can reconstruct the object root graph from it from my own process. It this cool or what? You can also peek into the Finalization Queue to check if you did accidentally forget to dispose a whole bunch of objects … /// <summary> /// .NET 4.0 or above only. Get all finalizable objects which are ready for finalization and have no other object roots anymore. /// </summary> static void NotYetFinalizedObjects() { using (var dumper = new MemoryDumper()) { object[] finalizable = dumper.GetObjectsReadyForFinalization(); Console.WriteLine("Currently {0} objects of types {1} are ready for finalization. Consider disposing them before.", finalizable.Length, String.Join(",", finalizable.ToLookup( x=> x.GetType() ) .Select( x=> x.Key.Name)) ); } } How does it work? The W of WMemoryProfiler is a good hint. It does employ Windbg and SOS dll to do the heavy lifting and concentrates on an easy to use Api which does hide completely Windbg. If you do not want to see Windbg you will never see it. In my experience the most complex thing is actually to download Windbg from the Windows 8 Stanalone SDK. This is described in the Readme and the exception you are greeted with if it is missing in much greater detail. So I will not go into this here.   What Next? Depending on the feedback I do get I can imagine some features which might be useful as well Calculate first order GC Roots from the actual object graph Identify global statics in Types in object graph Support read out of finalization queue of .NET 2.0 as well. Support Memory Dump analysis (again a feature only supported by commercial profilers in their professional editions if it is supported at all) Deserialize objects from a memory dump into a live process back (this would need some more investigation but it is doable) The last item needs some explanation. Why on earth would you want to do that? The basic idea is to store in your live process some logging/tracing data which can become quite big but since it is never written to it is very fast to generate. When your process crashes with a memory dump you could transfer this data structure back into a live viewer which can then nicely display your program state at the point it did crash. This is an advanced trouble shooting technique I have not seen anywhere yet but it could be quite useful. You can have here a look at the current feature list of WMemoryProfiler with some examples.   How To Get Started? First I would download the released source package (it is tiny). And compile the complete project. Then you can compile the Example project (it has this name) and uncomment in the main method the scenario you want to check out. If you are greeted with an exception it is time to install the Windows 8 Standalone SDK which is described in great detail in the exception text. Thats it for the first round. I have seen something more limited in the Java world some years ago (now I cannot find the link anymore) but anyway. Now we have something much better.

    Read the article

  • Triangle Picking Picking Back faces

    - by Tangeleno
    I'm having a bit of trouble with 3D picking, at first I thought my ray was inaccurate but it turns out that the picking is happening on faces facing the camera and faces facing away from the camera which I'm currently culling. Here's my ray creation code, I'm pretty sure the problem isn't here but I've been wrong before. private uint Pick() { Ray cursorRay = CalculateCursorRay(); Vector3? point = Control.Mesh.RayCast(cursorRay); if (point != null) { Tile hitTile = Control.TileMesh.GetTileAtPoint(point); return hitTile == null ? uint.MaxValue : (uint)(hitTile.X + hitTile.Y * Control.Generator.TilesWide); } return uint.MaxValue; } private Ray CalculateCursorRay() { Vector3 nearPoint = Control.Camera.Unproject(new Vector3(Cursor.Position.X, Control.ClientRectangle.Height - Cursor.Position.Y, 0f)); Vector3 farPoint = Control.Camera.Unproject(new Vector3(Cursor.Position.X, Control.ClientRectangle.Height - Cursor.Position.Y, 1f)); Vector3 direction = farPoint - nearPoint; direction.Normalize(); return new Ray(nearPoint, direction); } public Vector3 Camera.Unproject(Vector3 source) { Vector4 result; result.X = (source.X - _control.ClientRectangle.X) * 2 / _control.ClientRectangle.Width - 1; result.Y = (source.Y - _control.ClientRectangle.Y) * 2 / _control.ClientRectangle.Height - 1; result.Z = source.Z - 1; if (_farPlane - 1 == 0) result.Z = 0; else result.Z = result.Z / (_farPlane - 1); result.W = 1f; result = Vector4.Transform(result, Matrix4.Invert(ProjectionMatrix)); result = Vector4.Transform(result, Matrix4.Invert(ViewMatrix)); result = Vector4.Transform(result, Matrix4.Invert(_world)); result = Vector4.Divide(result, result.W); return new Vector3(result.X, result.Y, result.Z); } And my triangle intersection code. Ripped mainly from the XNA picking sample. public float? Intersects(Ray ray) { float? closestHit = Bounds.Intersects(ray); if (closestHit != null && Vertices.Length == 3) { Vector3 e1, e2; Vector3.Subtract(ref Vertices[1].Position, ref Vertices[0].Position, out e1); Vector3.Subtract(ref Vertices[2].Position, ref Vertices[0].Position, out e2); Vector3 directionCrossEdge2; Vector3.Cross(ref ray.Direction, ref e2, out directionCrossEdge2); float determinant; Vector3.Dot(ref e1, ref directionCrossEdge2, out determinant); if (determinant > -float.Epsilon && determinant < float.Epsilon) return null; float inverseDeterminant = 1.0f/determinant; Vector3 distanceVector; Vector3.Subtract(ref ray.Position, ref Vertices[0].Position, out distanceVector); float triangleU; Vector3.Dot(ref distanceVector, ref directionCrossEdge2, out triangleU); triangleU *= inverseDeterminant; if (triangleU < 0 || triangleU > 1) return null; Vector3 distanceCrossEdge1; Vector3.Cross(ref distanceVector, ref e1, out distanceCrossEdge1); float triangleV; Vector3.Dot(ref ray.Direction, ref distanceCrossEdge1, out triangleV); triangleV *= inverseDeterminant; if (triangleV < 0 || triangleU + triangleV > 1) return null; float rayDistance; Vector3.Dot(ref e2, ref distanceCrossEdge1, out rayDistance); rayDistance *= inverseDeterminant; if (rayDistance < 0) return null; return rayDistance; } return closestHit; } I'll admit I don't fully understand all of the math behind the intersection and that is something I'm working on, but my understanding was that if rayDistance was less than 0 the face was facing away from the camera, and shouldn't be counted as a hit. So my question is, is there an issue with my intersection or ray creation code, or is there another check I need to perform to tell if the face is facing away from the camera, and if so any hints on what that check might contain would be appreciated.

    Read the article

  • Cloud Computing Forces Better Design Practices

    - by Herve Roggero
    Is cloud computing simply different than on premise development, or is cloud computing actually forcing you to create better applications than you normally would? In other words, is cloud computing merely imposing different design principles, or forcing better design principles?  A little while back I got into a discussion with a developer in which I was arguing that cloud computing, and specifically Windows Azure in his case, was forcing developers to adopt better design principles. His opinion was that cloud computing was not yielding better systems; just different systems. In this blog, I will argue that cloud computing does force developers to use better design practices, and hence better applications. So the first thing to define, of course, is the word “better”, in the context of application development. Looking at a few definitions online, better means “superior quality”. As it relates to this discussion then, I stipulate that cloud computing can yield higher quality applications in terms of scalability, everything else being equal. Before going further I need to also outline the difference between performance and scalability. Performance and scalability are two related concepts, but they don’t mean the same thing. Scalability is the measure of system performance given various loads. So when developers design for performance, they usually give higher priority to a given load and tend to optimize for the given load. When developers design for scalability, the actual performance at a given load is not as important; the ability to ensure reasonable performance regardless of the load becomes the objective. This can lead to very different design choices. For example, if your objective is to obtains the fastest response time possible for a service you are building, you may choose the implement a TCP connection that never closes until the client chooses to close the connection (in other words, a tightly coupled service from a connectivity standpoint), and on which a connection session is established for faster processing on the next request (like SQL Server or other database systems for example). If you objective is to scale, you may implement a service that answers to requests without keeping session state, so that server resources are released as quickly as possible, like a REST service for example. This alternate design would likely have a slower response time than the TCP service for any given load, but would continue to function at very large loads because of its inherently loosely coupled design. An example of a REST service is the NO-SQL implementation in the Microsoft cloud called Azure Tables. Now, back to cloud computing… Cloud computing is designed to help you scale your applications, specifically when you use Platform as a Service (PaaS) offerings. However it’s not automatic. You can design a tightly-coupled TCP service as discussed above, and as you can imagine, it probably won’t scale even if you place the service in the cloud because it isn’t using a connection pattern that will allow it to scale [note: I am not implying that all TCP systems do not scale; I am just illustrating the scalability concepts with an imaginary TCP service that isn’t designed to scale for the purpose of this discussion]. The other service, using REST, will have a better chance to scale because, by design, it minimizes resource consumption for individual requests and doesn’t tie a client connection to a specific endpoint (which means you can easily deploy this service to hundreds of machines without much trouble, as long as your pockets are deep enough). The TCP and REST services discussed above are both valid designs; the TCP service is faster and the REST service scales better. So is it fair to say that one service is fundamentally better than the other? No; not unless you need to scale. And if you don’t need to scale, then you don’t need the cloud in the first place. However, it is interesting to note that if you do need to scale, then a loosely coupled system becomes a better design because it can almost always scale better than a tightly-coupled system. And because most applications grow overtime, with an increasing user base, new functional requirements, increased data and so forth, most applications eventually do need to scale. So in my humble opinion, I conclude that a loosely coupled system is not just different than a tightly coupled system; it is a better design, because it will stand the test of time. And in my book, if a system stands the test of time better than another, it is of superior quality. Because cloud computing demands loosely coupled systems so that its underlying service architecture can be leveraged, developers ultimately have no choice but to design loosely coupled systems for the cloud. And because loosely coupled systems are better… … the cloud forces better design practices. My 2 cents.

    Read the article

  • Waiting for Windows 8: A Long, Hot Summer

    - by andrewbrust
    Microsoft has revealed some things about Windows 8, and revealed a part of the developer story for new Windows 8 “tailored,” “immersive” applications.  In retrospect, very little was shared.  The bit that was revealed to us is that those applications can be developed using a combination of HTML 5 and JavaScript.  Not much else was said, except that additional details would be revealed at Microsoft’s //Build/ conference in Anaheim, California in September. This has left a lot of people in suspense, and it seems that suspended state is going to last all summer.  The problem, of course, is that in the absence of hard information, people fill the void with Speculation, Rumor and Gloom.  That’s a bit like Fear, Uncertainty and Doubt, except that it’s self-imposed by the Microsoft community and not planted by Microsoft’s competitors. This is a less-than-perfect situation.  Not only is it causing developers to worry about the value of their skill sets, but I am already hearing from consulting shops that customers are getting nervous too and, in extreme cases, opting for non-Microsoft tools for their projects as a result.  I’m also hearing from dev tool ISVs that sales have suffered as a result. It’s quite possible that the customers moving off .NET wanted to do so anyway and it’s also possible that dev tool ISVs are suffering slower sales this year due a slowed rate of economic recovery. Without hard information, tend to people interpret things negatively.  Actually, that’s the major point in all of this. While there is multitude of opinions about what the Windows 8 development platform will look like once fully revealed, there is an emerging consensus around one thing: it sure would help if Microsoft revealed more of its strategy…just enough to quash absurd rumors, stabilize the .NET ecosystem and get people to stay calm. We’ve had some reassurances thus far: there will be a Windows desktop mode; we’ll still have Windows Explorer, we’ll still run Office, we’ll still have a task bar, and all the skills and tools we use now will still work there.  But with reassurances like that…people still feel insecure.  Because telling us that Windows 8 will have what is essentially a “classic” mode sure makes it sound like today’s skill sets will soon be “classic” too…and then maybe they’ll just become obsolete. Humans find change scary; it’s natural.  And when left alone with their fears – because no one is saying anything to dispel them – people can go from frightened to paranoid, and can start to viewing things in a downright conspiratorial light.  It would be great if Microsoft stepped into the void now and told us what is coming – especially because whatever they tell us is bound to be at least a little better than what people think they are going to hear. I don’t know what the announcements will be, but I do have it on authority, from a number of sources, that Microsoft isn’t gong to talk until //Build/.  That means no news until September September 13th.  Nothing until after Labor Day.  You get zippo until after the Back-to-School sales are done. What to do?  Try not to let the dark voices of gloom and doom fill your head.  Even in the absence of answers, we still have some important facts: The .NET developer community is huge. Microsoft’s customers have major investments in .NET, and in .NET skills. Political infighting in Redmond might make for irrational decisions, but ultimately public companies can’t just alienate their advocates and piss off their customers.  Spite doesn’t trump fiduciary responsibility. The computing device markets are changing, software is changing, software business models are changing and developers are changing.  Microsoft has to keep up. The HTML + JavaScript community is huge too, and it includes many of the “changed” developers. Public companies can’t ignore new markets nor the popular standards that can help them enter those new markets.  Loyalty doesn’t trump fiduciary responsibility either. If Microsoft can appeal to new developers, then it should. If Microsoft can keep catering to its existing developers and customers -- not just through legacy support, but also through empowering futures -- then it probably will. You don’t have to shove your old friends out into the rain to make room for new ones; you can bring those new constituents in under a bigger tent.  I hope Microsoft will enlarge the tent, and I have trouble imagining why it would not.

    Read the article

  • Battery life starts at 2:30 hrs (99%), but less than 1 minute later is only 1:30 hrs (99%)

    - by zondu
    After searching this and other forums, I haven't seen this same issue listed anywhere for Ubuntu 12. Prior to installing Ubuntu 12.10, my Netbook (Acer AspireOne D250, SATA HDD) was consistently getting 2:30-3 hrs battery life under Windows XP Home, SP3. However, immediately after installing Ubuntu 12.10, the battery life starts out at 2:30 hrs (99%), but less than 1 minute later suddenly drops to 1:30 hrs (99%), which seems very odd. It could be a complete coincidence that the battery is suddenly flaky at the exact same moment that Ubuntu 12.10 was installed, but that doesn't seem likely. I'm a newbie to Ubuntu, so I don't have much experience tweaking/trouble-shooting yet. Here's what I've tried so far: enabled laptop mode (sudo su, then echo 5 /proc/sys/vm/laptop_mode) and checked that it is running when the A/C adapter is unplugged, but it doesn't seem to have made any noticeable difference in battery life, installed Jupiter, but it didn't work and messed up the system, so I had to uninstall it, disabled bluetooth (wifi is still on b/c it is necessary), set the screen to lowest brightness, etc., run through at least 1 full power cycle (running until the netbook shut itself off due to critical battery) and have been using it normally (sometimes plugged in, often unplugged until the battery gets very low) for a week since installing Ubuntu 12.10. installed powertop, but have no idea how to interpret its results. Here are the results of acpi -b: w/ A/C adapter: Battery 0: Full, 100% immediately after unplugging: Battery 0: Discharging, 99%, 02:30:20 remaining 1 minute after unplugging: Battery 0: Discharging, 99%, 01:37:49 remaining 2-3 minutes after unplugging: Battery 0: Discharging, 95%, 01:33:01 remaining 10 minutes after unplugging: Battery 0: Discharging, 85%, 01:13:38 remaining Results of cat /sys/class/power_supply/BAT0/uevent: w/ A/C adapter: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Full POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=12136000 POWER_SUPPLY_CURRENT_NOW=773000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1956000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= immediately after unplugging: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11886000 POWER_SUPPLY_CURRENT_NOW=773000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1937000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 1 minute later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11728000 POWER_SUPPLY_CURRENT_NOW=1174000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1937000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 2-3 minutes later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11583000 POWER_SUPPLY_CURRENT_NOW=1209000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1878000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 10 minutes later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11230000 POWER_SUPPLY_CURRENT_NOW=1239000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1644000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= Results of upower -i /org/freedesktop/UPower/devices/battery_BAT0: w/ A/C adapter: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:24:58 2012 (823 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: fully-charged energy: 21.1248 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 8.3484 W voltage: 12.173 V percentage: 100% capacity: 43.4667% technology: lithium-ion immediately after unplugging: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:41:25 2012 (1 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.9196 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 8.3484 W voltage: 11.86 V time to empty: 2.5 hours percentage: 99.0286% capacity: 43.4667% technology: lithium-ion History (charge): 1354023683 99.029 discharging 1 minute later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:42:31 2012 (17 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.9196 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.5432 W voltage: 11.753 V time to empty: 1.5 hours percentage: 99.0286% capacity: 43.4667% technology: lithium-ion History (charge): 1354023683 99.029 discharging History (rate): 1354023751 13.543 discharging 2-3 minutes later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:45:06 2012 (20 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.2824 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.7484 W voltage: 11.545 V time to empty: 1.5 hours percentage: 96.0123% capacity: 43.4667% technology: lithium-ion History (charge): 1354023906 96.012 discharging 1354023844 97.035 discharging History (rate): 1354023906 13.748 discharging 1354023875 12.992 discharging 1354023844 13.284 discharging 10 minutes later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:54:24 2012 (28 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 18.1764 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.2948 W voltage: 11.268 V time to empty: 1.4 hours percentage: 86.0429% capacity: 43.4667% technology: lithium-ion History (charge): 1354024433 86.043 discharging History (rate): 1354024464 13.295 discharging 1354024433 13.662 discharging 1354024402 13.781 discharging I noticed that between #2 and #3 (0 and 1 minutes after unplugging), while the battery still reports 99% charge and drops from 2:30 hr to 1:30 hr, the energy usage goes from 8.34 W to 13.54 W and the current_now increases, but shouldn't it be using less energy in battery mode since the screen is much dimmer and it's in power saving mode? (or is that normal behavior?) It also seems to drain more quickly than what it predicts, especially with the 1-1.25 hour drop in the first minute of being unplugged, which seems odd. What really concerns me is that Ubuntu 12.10 may not be properly managing the battery (with the sudden change in charge/life from 2:30 to 1:30 or 1:15 within a minute of unplugging), and that a new battery may quickly die under Ubuntu 12.10. I'd greatly appreciate any advice/suggestions on what to do, and especially whether there's a way to get back the 1-1.5 hrs of battery life that were suddenly lost when changing from WinXp to Ubuntu 12.10. Thanks :)

    Read the article

  • Styling ASP.NET MVC Error Messages

    - by MightyZot
    Originally posted on: http://geekswithblogs.net/MightyZot/archive/2013/11/11/styling-asp.net-mvc-error-messages.aspxOff the cuff, it may look like you’re stuck with the presentation of your error messages (model errors) in ASP.NET MVC. That’s not the case, though. You actually have quite a number of options with regard to styling those boogers. Like many of the helpers in MVC, the Html.ValidationMessageFor helper has multiple prototypes. One of those prototypes lets you pass a dictionary, or anonymous object, representing attribute values for the resulting markup. @Html.ValidationMessageFor( m => Model.Whatever, null, new { @class = “my-error” }) By passing the htmlAttributes parameter, which is the last parameter in the call to the prototype of Html.ValidationMessageFor shown above, I can style the resulting markup by associating styles to the my-error css class.  When you run your MVC project and view the source, you’ll notice that MVC adds the class field-validation-valid or field-validation-error to a span created by the helper. You could actually just style those classes instead of adding your own…it’s really up to you. Now, what if you wanted to move that error message around? Maybe you want to put that error message in a box or a callout. How do you do that? When I first started using MVC, it didn’t occur to me that the Html.ValidationMessageFor helper just spits out a little bit of markup. I wanted to put the error messages in boxes with white backgrounds, our site originally had a black background, and show a little nib on the side to make them look like callouts or conversation bubbles. Not realizing how much freedom there is in the styling and markup, and after reading someone else’s post, I created my own version of the ValidationMessageFor helper that took out the span and replaced it with divs. I styled the divs to produce the effect of a popup box and had a lot of trouble with sizing and such. That’s a really silly and unnecessary way to solve this problem. If you want to move your error messages around, all you have to do is move the helper. MVC doesn’t appear to care where you put it, which makes total sense when you think about it. Html.ValidationMessageFor is just spitting out a little markup using a little bit of reflection on the name you’re passing it. All you’ve got to do to style it the way you want it is to put it in whatever markup you desire. Take a look at this, for example… <div class=”my-anchor”>@Html.ValidationMessageFor( m => Model.Whatever )</div> @Html.TextBoxFor(m => Model.Whatever) Now, given that bit of HTML, consider the following CSS… <style> .my-anchor { position:relative; } .field-validation-error {    background-color:white;    border-radius:4px;    border: solid 1px #333;    display: block;    position: absolute;    top:0; right:0; left:0;    text-align:right; } </style> The my-anchor class establishes an anchor for the absolutely positioned error message. Now you can move the error message wherever you want it relative to the anchor. Using css3, there are some other tricks. For example, you can use the :not(:empty) selector to select the span and apply styles based upon whether or not the span has text in it. Keep it simple, though. Moving your elements around using absolute positioning may cause you issues on devices with screens smaller than your standard laptop or PC. While looking for something else recently, I saw someone asking how to style the output for Html.ValidationSummary.  Html.ValidationSummery is the helper that will spit out a list of property errors, general model errors, or both. Html.ValidationSummary spits out fairly simple markup as well, so you can use the techniques described above with it also. The resulting markup is a <ul><li></li></ul> unordered list of error messages that carries the class validation-summary-errors In the forum question, the user was asking how to hide the error summary when there are no errors. Their errors were in a red box and they didn’t want to show an empty red box when there aren’t any errors. Obviously, you can use the css3 selectors to apply different styles to the list when it’s empty and when it’s not empty; however, that’s not support in all browsers. Well, it just so happens that the unordered list carries the style validation-summary-valid when the list is empty. While the div rendered by the Html.ValidationSummary helper renders a visible div, containing one invisible listitem, you can always just style the whole div with “display:none” when the validation-summary-valid class is applied and make it visible when the validation-summary-errors class is applied. Or, if you don’t like that solution, which I like quite well, you can also check the model state for errors with something like this… int errors = ViewData.ModelState.Sum(ms => ms.Value.Errors.Count); That’ll give you a count of the errors that have been added to ModelState. You can check that and conditionally include markup in your page if you want to. The choice is yours. Obviously, doing most everything you can with styles increases the flexibility of the presentation of your solution, so I recommend going that route when you can. That picture of the fat guy jumping has nothing to do with the article. That’s just a picture of me on the roof and I thought it was funny. Doesn’t every post need a picture?

    Read the article

  • How to fix Ubuntu 12.04.3 boot to black screen full of errors in white text, after upgrading on dell inspiron 1501

    - by Ibuntu
    I am running a Dell Inspiron 1501 I use Linux only. No Microsoft or Apple operating systems (or really anything closed-source). I've only been using Linux for a little over a year but I'm starting to gain a comfortable level of familiarity with the system and terminology. I've been having some issues with Quantel Quetzal and Raring Ringtail, especially with older hardware, so I opted to install Ubuntu 12.04.3 Precise Pangolin on the Inspiron 1501. I checked my MD5 sum after downloading my ISO and all was good. I have in fact used this iso/dvd to install Precise Pangolin successfully on a few other systems (some of which are even older than this laptop). Install goes fine. The wireless card doesn't work out of the box but this is a known issue which is fairly easy to fix. So, first thing I did was open up a terminal and run sudo apt-get update && sudo apt-get upgrade which, part way through, crashed (I assume lightdm and possibly X) and took me to a black screen filled with white lines of text that were either errors or just the ouputs of commands. The reason I say that is because I was unable to gleam any useful information from the output on the screen. I did take a picture however and will post a link. After that, every time I boot the system it goes right to that black screen posting all the error messages or output in white text. I never get a purple Ubuntu splash, so from what I can tell after reading this wiki article: https://wiki.ubuntu.com/X/Troubleshooting/BlankScreen That means that after the kernel is selected, it is unable to correctly implement the settings it needs. If the purple splash never shows, the frame buffer was never set correctly right? This leads me to believe that it could be a kernel issue? The wiki suggested to try and pinpoint the issue by rolling back kernels until I find one that works. Is this my best option? I think I'm going to give it a try anyways and will let everyone know if I am able to solve the issue this way. I have since done a few reinstalls and some trouble-shooting including a couple hours scouring the net for anyone with any kind of similar issue. Most of the issues I could find involved getting a black screen after login and none of them said anything about any information output on this black screen. My reinstalls have taught me that there is no issue updating, but as soon as I run sudo apt-get upgrade my system goes to the black screen and every time I boot it up it does the same thing. The only way to fix is by reinstall. I never get any ability to log in. After a hard power off to the laptop (because I cannot use ctrl+alt+del to reboot) when it boots again it goes to the grub boot menu and I can select between regular boot, recovery mode and the two memtest options. I never tried the memtest options but the other two both lead to the same black screen. Some people having a black/blank screen issue claim to have fixed it by using 12.10 or 13.04 but I believe they were having a different issue where they got a black/blank screen after logging in. I think I will still give these images a try, but mostly figured I would just wait another day or two for 13.10. Other things I figured I would try from the following three articles: After logging in, there's a black screen and my cursor, nothing else! in Ubuntu 12.10 Black Screen on Login After Upgrading to 12.04 I can't get to the login screen include opening a terminal using ctrl+alt+f1 and trying a variety of reseting unity, x settings, lightdm (or switching to gdm); but I doubt this will work or that I will even be able to access a terminal. I'm pretty sure the whole system is stuck after it loads the last line on the black screen. I will try these things and post more information when I have. Hopefully someone has an idea in the meantime and I will keep checking back trying to find a solution. Thank you. Here are 3 different pictures of the error message. I had to take with my phone: http://ubuntuone.com/album/0TBBkxmVajJIQQtoN9mVdN

    Read the article

  • What Counts for A DBA: Observant

    - by drsql
    When walking up to the building where I work, I can see CCTV cameras placed here and there for monitoring access to the building. We are required to wear authorization badges which could be checked at any time. Do we have enemies?  Of course! No one is 100% safe; even if your life is a fairy tale, there is always a witch with an apple waiting to snack you into a thousand years of slumber (or at least so I recollect from elementary school.) Even Little Bo Peep had to keep a wary lookout.    We nerdy types (or maybe it was just me?) generally learned on the school playground to keep an eye open for unprovoked attack from simpler, but more muscular souls, and take steps to avoid messy confrontations well in advance. After we’d apprehensively negotiated adulthood with varying degrees of success, these skills of watching for danger, and avoiding it,  translated quite well to the technical careers so many of us were destined for. And nowhere else is this talent for watching out for irrational malevolence so appropriate as in a career as a production DBA.   It isn’t always active malevolence that the DBA needs to watch out for, but the even scarier quirks of common humanity.  A large number of the issues that occur in the enterprise happen just randomly or even just one time ever in a spurious manner, like in the case where a person decided to download the entire MSDN library of software, cross join every non-indexed billion row table together, and simultaneously stream the HD feed of 5 different sporting events, making the network access slow while the corporate online sales just started. The decent DBA team, like the going, gets tough under such circumstances. They spring into action, checking all of the sources of active information, observes the issue is no longer happening now, figures that either it wasn’t the database’s fault and that the reboot of the whatever device on the network fixed the problem.  This sort of reactive support is good, and will be the initial reaction of even excellent DBAs, but it is not the end of the story if you really want to know what happened and avoid getting called again when it isn’t even your fault.   When fires start raging within the corporate software forest, the DBA’s instinct is to actively find a way to douse the flames and get back to having no one in the company have any idea who they are.  Even better for them is to find a way of killing a potential problem while the fires are small, long before they can be classified as raging. The observant DBA will have already been monitoring the server environment for months in advance.  Most troubles, such as disk space and security intrusions, can be predicted and dealt with by alerting systems, whereas other trouble can come out of the blue and requires a skill of observing ongoing conditions and noticing inexplicable changes that could signal an emerging problem.  You can’t automate the DBA, because the bankable skill of a DBA is in detecting the early signs of unexpected problems, and working out how to deal with them before anyone else notices them.    To achieve this, the DBA will check the situation as it is currently happening,  and in many cases is likely to have been the person who submitted the problem to the level 1 support person in the first place, just to let the support team know of impending issues (always well received, I tell you what!). Database and host computer settings, configurations, and even critical data might be profiled and captured for later comparisons. He’ll use Monitoring tools, built-in, commercial (Not to be too crassly commercial or anything, but there is one such tool is SQL Monitor) and lots of homebrew monitoring tools to monitor for problems and changes in the server environment.   You will know that you have it right when a support call comes in and you can look at your monitoring tools and quickly respond that “response time is well within the normal range, the query that supports the failing interface works perfectly and has actually only been called 67% as often as normal, so I am more than willing to help diagnose the problem, but it isn’t the database server’s fault and is probably a client or networking slowdown causing the interface to be used less frequently than normal.” And that is the best thing for any DBA to observe…

    Read the article

  • Stumbling Through: Visual Studio 2010 (Part II)

    I would now like to expand a little on what I stumbled through in part I of my Visual Studio 2010 post and touch on a few other features of VS 2010.  Specifically, I want to generate some code based off of an Entity Framework model and tie it up to an actual data source.  Im not going to take the easy way and tie to a SQL Server data source, though, I will tie it to an XML data file instead.  Why?  Well, why not?  This is purely for learning, there are probably much better ways to get strongly-typed classes around XML but it will force us to go down a path less travelled and maybe learn a few things along the way.  Once we get this XML data and the means to interact with it, I will revisit data binding to this data in a WPF form and see if I cant get reading, adding, deleting, and updating working smoothly with minimal code.  To begin, I will use what was learned in the first part of this blog topic and draw out a data model for the MFL (My Football League) - I dont want the NFL to come down and sue me for using their name in this totally football-related article.  The data model looks as follows, with Teams having Players, and Players having a position and statistics for each season they played: Note that when making the associations between these entities, I was given the option to create the foreign key but I only chose to select this option for the association between Player and Position.  The reason for this is that I am picturing the XML that will contain this data to look somewhat like this: <MFL> <Position/> <Position/> <Position/> <Team>     <Player>         <Statistic/>     </Player> </Team> </MFL> Statistic will be under its associated Player node, and Player will be under its associated Team node no need to have an Id to reference it if we know it will always fall under its parent.  Position, however, is more of a lookup value that will not have any hierarchical relationship to the player.  In fact, the Position data itself may be in a completely different xml file (something Id like to play around with), so in any case, a player will need to reference the position by its Id. So now that we have a simple data model laid out, I would like to generate two things based on it:  A class for each entity with properties corresponding to each entity property An IO class with methods to get data for each entity, either all instances, by Id or by parent. Now my experience with code generation in the past has consisted of writing up little apps that use the code dom directly to regenerate code on demand (or using tools like CodeSmith).  Surely, there has got to be a more fun way to do this given that we are using the Entity Framework which already has built-in code generation for SQL Server support.  Lets start with that built-in stuff to give us a base to work off of.  Right click anywhere in the canvas of our model and select Add Code Generation Item: So just adding that code item seemed to do quite a bit towards what I was intending: It apparently generated a class for each entity, but also a whole ton more.  I mean a TON more.  Way too much complicated code was generated now that code is likely to be a black box anyway so it shouldnt matter, but we need to understand how to make this work the way we want it to work, so lets get ready to do some stumbling through that text template (tt) file. When I open the .tt file that was generated, right off the bat I realize there is going to be trouble there is no color coding, no intellisense no nothing!  That is going to make stumbling through more like groping blindly in the dark while handcuffed and hopping on one foot, which was one of the alternate titles I was considering for this blog.  Thankfully, the community comes to my rescue and I wont have to cast my mind back to the glory days of coding in VI (look it up, kids).  Using the Extension Manager (Available under the Tools menu), I did a quick search for tt editor in the Online Gallery and quickly found the Tangible T4 Editor: Downloading and installing this was a breeze, and after doing so I got some color coding and intellisense while editing the tt files.  If you will be doing any customizing of tt files, I highly recommend installing this extension.  Next, well see if that is enough help for us to tweak that tt file to do the kind of code generation that we wantDid you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

    Read the article

  • Why is my animation getting aborted?

    - by Homer_Simpson
    I have a class named Animation which handles my animations. The animation class can be called from multiple other classes. For example, the class Player.cs can call the animation class like this: Animation Playeranimation; Playeranimation = new Animation(TimeSpan.FromSeconds(2.5f), 80, 40, Animation.Sequences.forwards, 0, 5, false, true); //updating the animation public void Update(GameTime gametime) { Playeranimation.Update(gametime); } //drawing the animation public void Draw(SpriteBatch batch) { playeranimation.Draw(batch, PlayerAnimationSpritesheet, PosX, PosY, 0, SpriteEffects.None); } The class Lion.cs can call the animation class with the same code, only the animation parameters are changing because it's another animation that should be played: Animation Lionanimation; Lionanimation = new Animation(TimeSpan.FromSeconds(2.5f), 100, 60, Animation.Sequences.forwards, 0, 8, false, true); Other classes can call the animation class with the same code like the Player class. But sometimes I have some trouble with the animations. If an animation is running and then shortly afterwards another class calls the animation class too, the second animation starts but the first animation is getting aborted. In this case, the first animation couldn't run until it's end because another class started a new instance of the animation class. Why is an animation sometimes getting aborted when another animation starts? How can I solve this problem? My animation class: public class Animation { private int _animIndex, framewidth, frameheight, start, end; private TimeSpan PassedTime; private List<Rectangle> SourceRects = new List<Rectangle>(); private TimeSpan Duration; private Sequences Sequence; public bool Remove; private bool DeleteAfterOneIteration; public enum Sequences { forwards, backwards, forwards_backwards, backwards_forwards } private void forwards() { for (int i = start; i < end; i++) SourceRects.Add(new Rectangle(i * framewidth, 0, framewidth, frameheight)); } private void backwards() { for (int i = start; i < end; i++) SourceRects.Add(new Rectangle((end - 1 - i) * framewidth, 0, framewidth, frameheight)); } private void forwards_backwards() { for (int i = start; i < end - 1; i++) SourceRects.Add(new Rectangle(i * framewidth, 0, framewidth, frameheight)); for (int i = start; i < end; i++) SourceRects.Add(new Rectangle((end - 1 - i) * framewidth, 0, framewidth, frameheight)); } private void backwards_forwards() { for (int i = start; i < end - 1; i++) SourceRects.Add(new Rectangle((end - 1 - i) * framewidth, 0, framewidth, frameheight)); for (int i = start; i < end; i++) SourceRects.Add(new Rectangle(i * framewidth, 0, framewidth, frameheight)); } public Animation(TimeSpan duration, int frame_width, int frame_height, Sequences sequences, int start_interval, int end_interval, bool remove, bool deleteafteroneiteration) { Remove = remove; DeleteAfterOneIteration = deleteafteroneiteration; framewidth = frame_width; frameheight = frame_height; start = start_interval; end = end_interval; switch (sequences) { case Sequences.forwards: { forwards(); break; } case Sequences.backwards: { backwards(); break; } case Sequences.forwards_backwards: { forwards_backwards(); break; } case Sequences.backwards_forwards: { backwards_forwards(); break; } } Duration = duration; Sequence = sequences; } public void Update(GameTime dt) { PassedTime += dt.ElapsedGameTime; if (PassedTime > Duration) { PassedTime -= Duration; } var percent = PassedTime.TotalSeconds / Duration.TotalSeconds; if (DeleteAfterOneIteration == true) { if (_animIndex >= SourceRects.Count) Remove = true; _animIndex = (int)Math.Round(percent * (SourceRects.Count)); } else { _animIndex = (int)Math.Round(percent * (SourceRects.Count - 1)); } } public void Draw(SpriteBatch batch, Texture2D Textures, float PositionX, float PositionY, float Rotation, SpriteEffects Flip) { if (DeleteAfterOneIteration == true) { if (_animIndex >= SourceRects.Count) return; } batch.Draw(Textures, new Rectangle((int)PositionX, (int)PositionY, framewidth, frameheight), SourceRects[_animIndex], Color.White, Rotation, new Vector2(framewidth / 2.0f, frameheight / 2.0f), Flip, 0f); } }

    Read the article

  • Long pause when accessing DFS namespace

    - by Matt
    We've recently migrated our Windows network to use DFS for shared files. DFS is working well, except for one annoying problem: users experience a significant delay when they try to access a DFS namespace that they have not accessed for some time. I have tried to troubleshoot the issue but have not had any success so far, and I was hoping someone here may have some pointers to help resolve the problem. Firstly, some background on our network: The network uses a Windows 2008 functional level Active Directory domain with two Windows 2008 DCs and two DNS servers (one on each of the DCs). The network is DNS only - no WINS. All computers are located at the same site and connected by Gigabit Ethernet. We have approximately 20 Domain-based DFS namespaces in Windows 2008 mode, and each DFS namespace has two Windows 2008 DFS namespace servers (the same two servers for all namespaces). All namespace servers are in FQDN mode and all folder targets are specified using their FQDN. All computers are up-to-date with Service Packs and patches. The actual folder targets (i.e. the SMB shares our DFS folders point to) are scattered across several file and application servers, all running Windows 2008 bar two application servers which run Windows 2003 R2, with no replication setup at all (e.g. all DFS folders currently only have one folder target). Some more detail on the problem: The namespace access delay is generally 1 - 10 seconds long and seems to occur when a particular computer has not accessed the requested namespace for approximately five minutes or more. For example, if the user has not accessed \\domain.name\namespace1\ for more than five minutes and attempts to access \\domain.name\namespace1\ via Windows Explorer, the Explorer window will freeze for 1 - 10 seconds before finally resuming and displaying the folders that exist in \\domain.name\namespace1. If they then close the Explorer window and attempt to access \\domain.name\namespace1\ again within five minutes the contents will be displayed almost instantly - if they wait longer than five minutes it will go through the 1 - 10 second pause again. Once "inside" the namespace everything is nice and snappy, it's just the initial connection to the namespace that is slow. The browsing delays seem to affect all variants of Windows that we use (Windows 2008 x64 SP2, Windows 2003 R2 x86 SP2, Windows XP Pro x86 SP3) - it is possibly a bit worse in Windows XP / 2003 than in Windows 2008, but I'm not sure if the difference isn't just psychological. Accessing the underlying folder targets directly exhibits no delay at all - i.e. if the SMB shares pointed to by DFS are accessed directly (bypassing DFS) then there is no pause. During trouble-shooting I noticed that the "Cache duration" for all of our DFS roots is set to 300 seconds - 5 minutes. Given that this is the same amount of time required to trigger the pause I assume that this caching is somehow related, although I am unsure exactly what is cached on the client and hence what needs to be looked up again after 5 minutes have elapsed. In trying to resolve the problem I have already tried / checked the following (without success): Run dcdiag on both Domain Controllers - no problems found Done some basic DNS server checks without finding any problems - I don't know how to check the DNS servers in detail, but I would add that the network is not exhibiting any other strange behavior that may point to a DNS problem Disabled Anti-virus on clients and servers Removing one of the namespace servers from a couple of namespaces - no difference So that's where I'm up to - and I'm out of ideas. Can anyone suggest what may be causing the delays and/or what I should be trying next?

    Read the article

  • Weird networking problem ( Linksys, Windows 7 )

    - by Rohit Nair
    Okay it's a bit tough to figure out where to start from, but here is the basic summary of the issue: During general internet usage, there are times when any attempt to visit a website stalls at "Waiting for somedomain.com". This problem occurs in Firefox, IE and Chrome. No website will load, INCLUDING the router configuration page at 192.168.1.1. Curiously, ping works fine, and other network apps such as MSN Messenger continue to work and I can send and receive messages. Disconnecting and reconnecting to the wireless network seems to fix the problem for a bit, but there are times when it relapses into not loading after every 2-3 http requests. Restarting the router seems to fix the issue, but it can crop up hours or days later. I have a CCNA cert and I know my way around the Windows family of operating systems, so I'm going to list all the things I've tried here. Other computers on the network seem to suffer the same problem, which makes me think it might be a specific problem with something in Win7. The random nature of this issue makes it a bit difficult to confirm, but I can definitely say that I have experienced this on the following systems: Windows 7 64-bit on my desktop Windows Vista 32-bit on my desktop ( the desktop has 2 wireless NICs and the problem existed on both ) Windows Vista 32-bit on my laptop ( both with wireless and wired ) Windows XP SP3 on another laptop ( both wireless and wired ) Using Wireshark to sniff packets seemed to indicate that although HTTP requests were being SENT out, no packets were coming in to respond to the HTTP request. However, other network apps continued to work i.e I would still receive IMs on Windows Live Messenger. Disabling IPV6 had no effect. Updating router firmware to the latest stock firmware by Linksys had no effect. Switching to dd-wrt firmware had no effect. By "no effect" I mean that although the restart required by firmware updates fixed the problem at the time, it still came back. A couple of weeks back, after a LOT of googling and flipping of various options, I figured it might be a case of router slowdown ( http://www.dd-wrt.com/wiki/index.php/Router%5FSlowdown ) caused by the fact that I occasionally run a torrent client. I tried changing the configuration as suggested in that router slowdown link, and restarted the router. However I have not run the torrent client for 12 days now, and yet I still randomly experience this problem. Currently the computer I am using is running Windows 7 64-bit. I would just like to reiterate some of the reasons that I was confused by the issue. Even the router config page at 192.168.1.1 would not load, indicating that it's not a problem with the WAN link, but probably a router issue or a local computer issue. For some reason, disconnecting and reconnecting to the wireless network immediately seems to fix the problem. Updating the router firmware, even switching to open source firmware did nothing. So it seemed to be a computer issue. On the other hand, I have not seen any mass outrage of people having networking problems with Windows 7 and Linksys routers, especially a problem of this sort, and I have tweaked every network setting I could think of. Although HTTP seems to have trouble, ping works fine, DNS lookups work fine, other networking apps work fine. However if I disconnect from Windows Live Messenger and try to reconnect, it fails to reconnect. So although it could receive data over the existing TCP/IP connection, trying to start a new one failed? Does anyone have any further ideas on debugging or fixing this issue? I am reasonably certain there are no viruses or other malicious apps on my network, and I am also reasonably certain that nobody is accessing my router without my consent. Router: Linksys WRT54G2 1.0 running dd-wrt firmware Wireless Card: Alfa AWUS036H OS: Windows 7 64-bit EDIT: I tried switching to a clean wireless channel free from interference, but the problem still persisted. I tried connecting directly with a cable, but the problem still persisted. Signed A very confused and bewildered geek whose knowledge seems to be useless in the face of this frustrating network issue.

    Read the article

  • Connecting debian and windows via IPsec VPN with Racoon and ipsec-tools

    - by Michi Qne
    I've some trouble with the IPsec configuration on my debian server (6 squeeze). This server should connect via IPsec VPN to an windows server, which is protected by an firewall. I've used racoon and ipsec-tools and this tutorial http://wiki.debian.org/IPsec. However, I am not quite sure, if this tutorial fits to my purpose, because of some differences: my Host and my gateway are the same server. So I don't have two different ip addresses. I guess, that's not a problem the other server is an windows system behind a firewall. Hopefully, not a problem the subnet of the windows system is /32 not /24. So I change it to /32. I worked through the tutorial step by step, but I wasn't able to route the ip. The following command didn't work for me: ip route add to 172.16.128.100/32 via XXX.XXX.XXX.XXX src XXX.XXX.XXX.XXX So I tried the following instead: ip route add to 172.16.128.100 .., which obviously not solved the problem. The next problem is the compression. The windows doesn't use a compression, but 'compression_algorithm none;' doesn't work with my racoon. So the current value is 'compression_algorithm deflate;' So my current result looks like this: When I am trying to ping the windows host (ping 172.16.128.100), I receive the following error message from ping: ping: sendmsg: Operation not permitted And racoon logs: racoon: ERROR: failed to get sainfo. After googling for a while I came to no conclusion, what's the solution. Does this error message mean that the first phase of IPsec works? I am thankful for any advice. I guess my configs might be helpful. My racoon.conf looks like this: path pre_shared_key "/etc/racoon/psk.txt"; remote YYY.YYY.YYY.YYY { exchange_mode main; proposal { lifetime time 8 hour; encryption_algorithm 3des; hash_algorithm sha1; authentication_method pre_shared_key; dh_group 2; } } sainfo address XXX.XXX.XXX.XXX/32 any address 172.16.128.100/32 any { pfs_group 2; lifetime time 8 hour; encryption_algorithm aes 256; authentication_algorithm hmac_sha1; compression_algorithm deflate; } And my ipsec-tools.conf looks like this: flush; spdflush; spdadd XXX.XXX.XXX.XXX/32 172.16.128.100/32 any -P out ipsec esp/tunnel/XXX.XXX.XXX.XXX-YYY.YYY.YYY.YYY/require; spdadd 172.16.128.100/32 XXX.XXX.XXX.XXX/32 any -P in ipsec esp/tunnel/YYY.YYY.YYY.YYY-XXX.XXX.XXX.XXX/require; If anyone has an advice, that would be awesome. Thanks in Advance. Greets, Michael It was a simple copy-and-paste error in an ip address.

    Read the article

  • Delaying NIS & NFS startup till after network interface is fully ready on Fedora 17

    - by obmarg
    I've recently set up a fedora 17 server for our network, and I've been having trouble getting the NIS service to work on startup. Here's some logs from the system: Aug 21 12:57:12 cairnwell ypbind-pre-setdomain[718]: Setting NIS domain: 'indigo-nis' (environment variable) Aug 21 12:57:13 cairnwell ypbind: Binding NIS service Aug 21 12:57:13 cairnwell rpc.statd[730]: Unable to prune capability 0 from bounding set: Operation not permitted Aug 21 12:57:13 cairnwell systemd[1]: nfs-lock.service: control process exited, code=exited status=1 Aug 21 12:57:13 cairnwell systemd[1]: Unit nfs-lock.service entered failed state. Aug 21 12:57:14 cairnwell setroubleshoot: SELinux is preventing /usr/sbin/rpc.statd from using the setpcap capability. For complete SELinux messages. run sealert -l 024bba8a-b0ef-43dc-b195-5c9a2d4c4d41 Aug 21 12:57:15 cairnwell kernel: [ 18.822282] bnx2 0000:02:00.0: em1: NIC Copper Link is Up, 1000 Mbps full duplex Aug 21 12:57:15 cairnwell kernel: [ 18.822925] ADDRCONF(NETDEV_CHANGE): em1: link becomes ready Aug 21 12:57:15 cairnwell NetworkManager[621]: <info> (em1): carrier now ON (device state 20) Aug 21 12:57:15 cairnwell NetworkManager[621]: <info> (em1): device state change: unavailable -> disconnected (reason 'carrier-changed') [20 30 40] Aug 21 12:57:15 cairnwell NetworkManager[621]: <info> Auto-activating connection 'System em1'. Aug 21 12:57:15 cairnwell NetworkManager[621]: <info> Activation (em1) starting connection 'System em1' Aug 21 12:57:15 cairnwell NetworkManager[621]: <info> (em1): device state change: disconnected -> prepare (reason 'none') [30 40 0] ....... Aug 21 12:57:19 cairnwell sendmail[790]: YPBINDPROC_DOMAIN: Domain not bound Aug 21 12:57:26 cairnwell sendmail[790]: YPBINDPROC_DOMAIN: Domain not bound Aug 21 12:57:31 cairnwell sendmail[790]: YPBINDPROC_DOMAIN: Domain not bound Aug 21 12:57:35 cairnwell sendmail[790]: YPBINDPROC_DOMAIN: Domain not bound Aug 21 12:58:00 cairnwell ypbind: Binding took 47 seconds Aug 21 12:58:00 cairnwell ypbind: NIS server for domain indigo-nis is not responding. Aug 21 12:58:01 cairnwell ypbind: Killing ypbind with PID 733. Aug 21 12:58:01 cairnwell ypbind-post-waitbind[734]: /usr/lib/ypbind/ypbind-post-waitbind: line 51: kill: SIGTERM: invalid signal specification Aug 21 12:58:01 cairnwell systemd[1]: ypbind.service: control process exited, code=exited status=1 Aug 21 12:58:01 cairnwell systemd[1]: Unit ypbind.service entered failed state. By the looks of these logs the ypbind service is starting up at 12:57:12 but the network interface isn't coming up till 12:57:15. My guess is that this is causing ypbind to time out when trying to connect. As a knock-on effect the NIS failure is causing problems with NFS which is no longer able to map UIDs properly. This problem doesn't seem to be fixed by actually starting ypbind etc. so I've had to set all my NFS shares to noauto. I have tried manually adding NETWORKDELAY and NETWORKWAIT in /etc/sysconfig/network and also running systemctl enable NetworkManager-wait-online.service as I've seen suggested in some places, but neither of these have had any effect. It is relatively easy to fix manually by restarting ypbind & mounting NFS shares after the network has started up, but it's less than ideal to have to do this every time the server has been rebooted. Does anyone know of an easy (and preferably hack-free) way of delaying the ypbind startup till after the network interface is fully ready?

    Read the article

  • iptables not allowing mysql connections to aliased ips?

    - by Curtis
    I have a fairly simple iptables firewall on a server that provides MySQL services, but iptables seems to be giving me very inconsistent results. The default policy on the script is as follows: iptables -P INPUT DROP I can then make MySQL public with the following rule: iptables -A INPUT -p tcp --dport 3306 -j ACCEPT With this rule in place, I can connect to MySQL from any source IP to any destination IP on the server without a problem. However, when I try to restrict access to just three IPs by replacing the above line with the following, I run into trouble (xxx=masked octect): iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.XXX.XXX.184 -j ACCEPT iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.XXX.XXX.196 -j ACCEPT iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.XXX.XXX.251 -j ACCEPT Once the above rules are in place, the following happens: I can connect to the MySQL server from the .184, .196 and .251 hosts just fine as long as am connecting to the MySQL server using it's default IP address or an IP alias in the same subnet as the default IP address. I am unable to connect to MySQL using IP aliases that are assigned to the server from a different subnet than the server's default IP when I'm coming from the .184 or .196 hosts, but .251 works just fine. From the .184 or .196 hosts, a telnet attempt just hangs... # telnet 209.xxx.xxx.22 3306 Trying 209.xxx.xxx.22... If I remove the .251 line (making .196 the last rule added), the .196 host still can not connect to MySQL using IP aliases (so it's not the order of the rules that is causing the inconsistent behavior). I know, this particular test was silly as it shouldn't matter what order these three rules are added in, but I figured someone might ask. If I switch back to the "public" rule, all hosts can connect to the MySQL server using either the default or aliased IPs (in either subnet): iptables -A INPUT -p tcp --dport 3306 -j ACCEPT The server is running in a CentOS 5.4 OpenVZ/Proxmox container (2.6.32-4-pve). And, just in case you prefer to see the problem rules in the context of the iptables script, here it is (xxx=masked octect): # Flush old rules, old custom tables /sbin/iptables --flush /sbin/iptables --delete-chain # Set default policies for all three default chains /sbin/iptables -P INPUT DROP /sbin/iptables -P FORWARD DROP /sbin/iptables -P OUTPUT ACCEPT # Enable free use of loopback interfaces /sbin/iptables -A INPUT -i lo -j ACCEPT /sbin/iptables -A OUTPUT -o lo -j ACCEPT # All TCP sessions should begin with SYN /sbin/iptables -A INPUT -p tcp ! --syn -m state --state NEW -j DROP # Accept inbound TCP packets (Do this *before* adding the 'blocked' chain) /sbin/iptables -A INPUT -m state --state ESTABLISHED,RELATED -j ACCEPT # Allow the server's own IP to connect to itself /sbin/iptables -A INPUT -i eth0 -s 208.xxx.xxx.178 -j ACCEPT # Add the 'blocked' chain *after* we've accepted established/related connections # so we remain efficient and only evaluate new/inbound connections /sbin/iptables -N BLOCKED /sbin/iptables -A INPUT -j BLOCKED # Accept inbound ICMP messages /sbin/iptables -A INPUT -p ICMP --icmp-type 8 -j ACCEPT /sbin/iptables -A INPUT -p ICMP --icmp-type 11 -j ACCEPT # ssh (private) /sbin/iptables -A INPUT -p tcp --dport 22 -m state --state NEW -s xxx.xxx.xxx.xxx -j ACCEPT # ftp (private) /sbin/iptables -A INPUT -p tcp --dport 21 -m state --state NEW -s xxx.xxx.xxx.xxx -j ACCEPT # www (public) /sbin/iptables -A INPUT -p tcp --dport 80 -j ACCEPT /sbin/iptables -A INPUT -p tcp --dport 443 -j ACCEPT # smtp (public) /sbin/iptables -A INPUT -p tcp --dport 25 -j ACCEPT /sbin/iptables -A INPUT -p tcp --dport 2525 -j ACCEPT # pop (public) /sbin/iptables -A INPUT -p tcp --dport 110 -j ACCEPT # mysql (private) /sbin/iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.xxx.xxx.184 -j ACCEPT /sbin/iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.xxx.xxx.196 -j ACCEPT /sbin/iptables -A INPUT -p tcp --dport 3306 -m state --state NEW -s 208.xxx.xxx.251 -j ACCEPT Any ideas? Thanks in advance. :-)

    Read the article

  • Error installing pkgconfig via macports

    - by Greg K
    I installed Macports 1.8.2 from a DMG. That seemed to install fine. I ran sudo port selfupdate to make sure my ports tree was current. I then tried to install bindfs as I want to mount some directories in my OS X file system (like you can do with mount --bind in linux). pkgconfig and macfuse are two dependencies of bindfs. I had trouble installing bindfs due to errors installing pkgconfig, so I tried to just install pkgconfig, here's the debug output from sudo port install pkgconfig: $ sudo port -d install pkgconfig DEBUG: Found port in file:///opt/local/var/macports/sources/rsync.macports.org/release/ports/devel/pkgconfig DEBUG: Changing to port directory: /opt/local/var/macports/sources/rsync.macports.org/release/ports/devel/pkgconfig DEBUG: OS Platform: darwin DEBUG: OS Version: 10.3.0 DEBUG: Mac OS X Version: 10.6 DEBUG: System Arch: i386 DEBUG: setting option os.universal_supported to yes DEBUG: org.macports.load registered provides 'load', a pre-existing procedure. Target override will not be provided DEBUG: org.macports.unload registered provides 'unload', a pre-existing procedure. Target override will not be provided DEBUG: org.macports.distfiles registered provides 'distfiles', a pre-existing procedure. Target override will not be provided DEBUG: adding the default universal variant DEBUG: Reading variant descriptions from /opt/local/var/macports/sources/rsync.macports.org/release/ports/_resources/port1.0/variant_descriptions.conf DEBUG: Requested variant darwin is not provided by port pkgconfig. DEBUG: Requested variant i386 is not provided by port pkgconfig. DEBUG: Requested variant macosx is not provided by port pkgconfig. ---> Computing dependencies for pkgconfig DEBUG: Executing org.macports.main (pkgconfig) DEBUG: Skipping completed org.macports.fetch (pkgconfig) DEBUG: Skipping completed org.macports.checksum (pkgconfig) DEBUG: Skipping completed org.macports.extract (pkgconfig) DEBUG: Skipping completed org.macports.patch (pkgconfig) ---> Configuring pkgconfig DEBUG: Using compiler 'Mac OS X gcc 4.2' DEBUG: Executing org.macports.configure (pkgconfig) DEBUG: Environment: CFLAGS='-O2 -arch x86_64' CPPFLAGS='-I/opt/local/include' CXXFLAGS='-O2 -arch x86_64' MACOSX_DEPLOYMENT_TARGET='10.6' CXX='/usr/bin/g++-4.2' F90FLAGS='-O2 -m64' LDFLAGS='-L/opt/local/lib' OBJC='/usr/bin/gcc-4.2' FCFLAGS='-O2 -m64' INSTALL='/usr/bin/install -c' OBJCFLAGS='-O2 -arch x86_64' FFLAGS='-O2 -m64' CC='/usr/bin/gcc-4.2' DEBUG: Assembled command: 'cd "/opt/local/var/macports/build/_opt_local_var_macports_sources_rsync.macports.org_release_ports_devel_pkgconfig/work/pkg-config-0.23" && ./configure --prefix=/opt/local --enable-indirect-deps --with-pc-path=/opt/local/lib/pkgconfig:/opt/local/share/pkgconfig' checking for a BSD-compatible install... /usr/bin/install -c checking whether build environment is sane... yes checking for gawk... no checking for mawk... no checking for nawk... no checking for awk... awk checking whether make sets $(MAKE)... no checking whether to enable maintainer-specific portions of Makefiles... no checking build system type... i386-apple-darwin10.3.0 checking host system type... i386-apple-darwin10.3.0 checking for style of include used by make... none checking for gcc... /usr/bin/gcc-4.2 checking for C compiler default output file name... configure: error: C compiler cannot create executables See `config.log' for more details. Error: Target org.macports.configure returned: configure failure: shell command " cd "/opt/local/var/macports/build/_opt_local_var_macports_sources_rsync.macports.org_release_ports_devel_pkgconfig/work/pkg-config-0.23" && ./configure --prefix=/opt/local --enable-indirect-deps --with-pc-path=/opt/local/lib/pkgconfig:/opt/local/share/pkgconfig " returned error 77 DEBUG: Backtrace: configure failure: shell command " cd "/opt/local/var/macports/build/_opt_local_var_macports_sources_rsync.macports.org_release_ports_devel_pkgconfig/work/pkg-config-0.23" && ./configure --prefix=/opt/local --enable-indirect-deps --with-pc-path=/opt/local/lib/pkgconfig:/opt/local/share/pkgconfig " returned error 77 while executing "$procedure $targetname" Warning: the following items did not execute (for pkgconfig): org.macports.activate org.macports.configure org.macports.build org.macports.destroot org.macports.install Error: Status 1 encountered during processing. I have only recently installed Xcode 3.2.2 (prior to installing macports). Am I right in thinking this the issue here: configure: error: C compiler cannot create executables

    Read the article

  • Rails, Capistrano, Nginx, Unicorn - Application has been already initialized (RuntimeError)

    - by Andy Copley
    Can anyone shed some light on what exactly this error refers to? I'm having trouble deploying new versions of the site. I, INFO -- : reloading config_file=[snip]/current/config/unicorn.rb I, INFO -- : Refreshing Gem list E, ERROR -- : error reloading config_file=[snip]/current/config/unicorn.rb: Application has been already initialized. (RuntimeError) E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/railties-3.2.3/lib/rails/application.rb:135:in `initialize!' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/railties-3.2.3/lib/rails/railtie/configurable.rb:30:in `method_missing' E, ERROR -- : [snip]/releases/20120907085937/config/environment.rb:5:in `<top (required)>' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/activesupport-3.2.3/lib/active_support/dependencies.rb:251:in `require' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/activesupport-3.2.3/lib/active_support/dependencies.rb:251:in `block in require' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/activesupport-3.2.3/lib/active_support/dependencies.rb:236:in `load_dependency' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/activesupport-3.2.3/lib/active_support/dependencies.rb:251:in `require' E, ERROR -- : config.ru:4:in `block in <main>' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/rack-1.4.1/lib/rack/builder.rb:51:in `instance_eval' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/rack-1.4.1/lib/rack/builder.rb:51:in `initialize' E, ERROR -- : config.ru:1:in `new' E, ERROR -- : config.ru:1:in `<main>' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn.rb:44:in `eval' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn.rb:44:in `block in builder' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn/http_server.rb:696:in `call' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn/http_server.rb:696:in `build_app!' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn/http_server.rb:677:in `load_config!' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/lib/unicorn/http_server.rb:303:in `join' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/gems/unicorn-4.3.1/bin/unicorn:121:in `<top (required)>' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/bin/unicorn:23:in `load' E, ERROR -- : [snip]/shared/bundle/ruby/1.9.1/bin/unicorn:23:in `<main>' I, INFO -- : reaped #<Process::Status: pid 3182 exit 0> worker=0 I, INFO -- : reaped #<Process::Status: pid 3185 exit 0> worker=1 I, INFO -- : reaped #<Process::Status: pid 3188 exit 0> worker=2 I, INFO -- : reaped #<Process::Status: pid 3191 exit 0> worker=3 I, INFO -- : worker=0 ready I, INFO -- : worker=3 ready I, INFO -- : worker=1 ready I, INFO -- : worker=2 ready Unicorn.rb root = "/home/[user]/apps/[site]/current" working_directory root pid "#{root}/tmp/pids/unicorn.pid" stderr_path "#{root}/log/unicorn.log" stdout_path "#{root}/log/unicorn.log" listen "/tmp/unicorn.[site].sock", :backlog => 2048 worker_processes 4 preload_app true timeout 30 before_fork do |server, worker| defined?(ActiveRecord::Base) and ActiveRecord::Base.connection.disconnect! end after_fork do |server, worker| defined?(ActiveRecord::Base) and ActiveRecord::Base.establish_connection end Any help appreciated - I can pull more config files if required.

    Read the article

  • Cannot join Win7 workstations to Win2k8 domain

    - by wfaulk
    I am trying to connect a Windows 7 Ultimate machine to a Windows 2k8 domain and it's not working. I get this error: Note: This information is intended for a network administrator. If you are not your network's administrator, notify the administrator that you received this information, which has been recorded in the file C:\Windows\debug\dcdiag.txt. DNS was successfully queried for the service location (SRV) resource record used to locate a domain controller for domain "example.local": The query was for the SRV record for _ldap._tcp.dc._msdcs.example.local The following domain controllers were identified by the query: dc1.example.local dc2.example.local However no domain controllers could be contacted. Common causes of this error include: Host (A) or (AAAA) records that map the names of the domain controllers to their IP addresses are missing or contain incorrect addresses. Domain controllers registered in DNS are not connected to the network or are not running. The client is in an office connected remotely via MPLS to the data center where our domain controllers exist. I don't seem to have anything blocking connectivity to the DCs, but I don't have total control over the MPLS circuit, so it's possible that there's something blocking connectivity. I have tried multiple clients (Win7 Ultimate and WinXP SP3) in the one office and get the same symptoms on all of them. I have no trouble connecting to either of the domain controllers, though I have, admittedly, not tried every possible port. ICMP, LDAP, DNS, and SMB connections all work fine. Client DNS is pointing to the DCs, and "example.local" resolves to the two IP addresses of the DCs. I get this output from the NetLogon Test command line utility: C:\Windows\System32>nltest /dsgetdc:example.local Getting DC name failed: Status = 1355 0x54b ERROR_NO_SUCH_DOMAIN I have also created a separate network to emulate that office's configuration that's connected to the DC network via LAN-to-LAN VPN instead of MPLS. Joining Windows 7 computers from that remote network works fine. The only difference I can find between the two environments is the intermediate connectivity, but I'm out of ideas as to what to test or how to do it. What further steps should I take? (Note that this isn't actually my client workstation and I have no direct access to it; I'm forced to do remote hands access to it, which makes some of the obvious troubleshooting methods, like packet sniffing, more difficult. If I could just set up a system there that I could remote into, I would, but requests to that effect have gone unanswered.) 2011-08-25 update: I had DCDIAG.EXE run on a client attempting to join the domain: C:\Windows\System32>dcdiag /u:example\adminuser /p:********* /s:dc2.example.local Directory Server Diagnosis Performing initial setup: Ldap search capabality attribute search failed on server dc2.example.local, return value = 81 This sounds like it was able to connect via LDAP, but the thing that it was trying to do failed. But I don't quite follow what it was trying to do, much less how to reproduce it or resolve it. 2011-08-26 update: Using LDP.EXE to try and make an LDAP connection directly to the DCs results in these errors: ld = ldap_open("10.0.0.1", 389); Error <0x51: Fail to connect to 10.0.0.1. ld = ldap_open("10.0.0.2", 389); Error <0x51: Fail to connect to 10.0.0.2. ld = ldap_open("10.0.0.1", 3268); Error <0x51: Fail to connect to 10.0.0.1. ld = ldap_open("10.0.0.2", 3268); Error <0x51: Fail to connect to 10.0.0.2. This would seem to point fingers at LDAP connections being blocked somewhere. (And 0x51 == 81, which was the error from DCDIAG.EXE from yesterday's update.) I could swear I tested this using TELNET.EXE weeks ago, but now I'm thinking that I may have assumed that its clearing of the screen was telling me that it was waiting and not that it had connected. I'm tracking down LDAP connectivity problems now. This update may become an answer.

    Read the article

< Previous Page | 220 221 222 223 224 225 226 227 228 229 230 231  | Next Page >