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• How to handle multi-processing of libraries which already spawn sub-processes?

  - by exhuma

  I am having some trouble coming up with a good solution to limit sub-processes in a script which uses a multi-processed library and the script itself is also multi-processed. Both, the library and script are modifiable by us. I believe the question is more about design than actual code, but for what it's worth, it's written in Python. The goal of the library is to hide implementation details of various internet routers. For that reason, the library has a "Proxy" factory method which takes the IP of a router as parameter. The factory then probes the device using a set of possible proxies. Usually, there is one proxy which immediately knows that is is able to send commands to this device. All others usually take some time to return (given a timeout). One thought was already to simply query the device for an identifier, and then select the proper proxy using that, but in order to do so, you would already need to know how to query the device. Abstracting this knowledge is one of the main purposes of the library, so that becomes a little bit of a "circular-requirement"/deadlock: To connect to a device, you need to know what proxy to use, and to know what proxy to create, you need to connect to a device. So probing the device is - as we can see - the best solution so far, apart from keeping a lookup-table somewhere. The library currently kills all remaining processes once a valid proxy has been found. And yes, there is always only one good proxy per device. Currently there are about 12 proxies. So if one create a proxy instance using the factory, 12 sub-processes are spawned. So far, this has been really useful and worked very well. But recently someone else wanted to use this library to "broadcast" a command to all devices. So he took the library, and wrote his own multi-processed script. This obviously spawned 12 * n processes where n is the number of IPs to which he broadcasted. This has given us two problems: The host on which the command was executed slowed down to a near halt. Aborting the script with CTRL+C ground the system to a total halt. Not even the hardware console responded anymore! This may be due to some Python strangeness which still needs to be investigated. Maybe related to http://bugs.python.org/issue8296 The big underlying question, is how to design a library which does multi-processing, so other applications which use this library and want to be multi-processed themselves do not run into system limitations. My first thought was to require a pool to be passed to the library, and execute all tasks in that pool. In that way, the person using the library has control over the usage of system resources. But my gut tells me that there must be a better solution. Disclaimer: My experience with multiprocessing is fairly limited. I have implemented a few straightforward which did not require access control to resources. So I have not yet any practical experience with semaphores or mutexes. p.s.: In the future, we may have enough information to do this without the probing. But the database which would contain the proper information is not yet operational. Also, the design about multiprocessing a multiprocessed library intrigues me :)

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• Validation and authorization in layered architecture

  - by SonOfPirate

  I know you are thinking (or maybe yelling), "not another question asking where validation belongs in a layered architecture?!?" Well, yes, but hopefully this will be a little bit of a different take on the subject. I am a firm believer that validation takes many forms, is context-based and varies at each level of the architecture. That is the basis for the post - helping to identify what type of validation should be performed in each layer. In addition, a question that often comes up is where authorization checks belong. The example scenario comes from an application for a catering business. Periodically during the day, a driver may turn in to the office any excess cash they've accumulated while taking the truck from site to site. The application allows a user to record the 'cash drop' by collecting the driver's ID, and the amount. Here's some skeleton code to illustrate the layers involved: public class CashDropApi // This is in the Service Facade Layer { [WebInvoke(Method = "POST")] public void AddCashDrop(NewCashDropContract contract) { // 1 Service.AddCashDrop(contract.Amount, contract.DriverId); } } public class CashDropService // This is the Application Service in the Domain Layer { public void AddCashDrop(Decimal amount, Int32 driverId) { // 2 CommandBus.Send(new AddCashDropCommand(amount, driverId)); } } internal class AddCashDropCommand // This is a command object in Domain Layer { public AddCashDropCommand(Decimal amount, Int32 driverId) { // 3 Amount = amount; DriverId = driverId; } public Decimal Amount { get; private set; } public Int32 DriverId { get; private set; } } internal class AddCashDropCommandHandler : IHandle<AddCashDropCommand> { internal ICashDropFactory Factory { get; set; } // Set by IoC container internal ICashDropRepository CashDrops { get; set; } // Set by IoC container internal IEmployeeRepository Employees { get; set; } // Set by IoC container public void Handle(AddCashDropCommand command) { // 4 var driver = Employees.GetById(command.DriverId); // 5 var authorizedBy = CurrentUser as Employee; // 6 var cashDrop = Factory.CreateCashDrop(command.Amount, driver, authorizedBy); // 7 CashDrops.Add(cashDrop); } } public class CashDropFactory { public CashDrop CreateCashDrop(Decimal amount, Employee driver, Employee authorizedBy) { // 8 return new CashDrop(amount, driver, authorizedBy, DateTime.Now); } } public class CashDrop // The domain object (entity) { public CashDrop(Decimal amount, Employee driver, Employee authorizedBy, DateTime at) { // 9 ... } } public class CashDropRepository // The implementation is in the Data Access Layer { public void Add(CashDrop item) { // 10 ... } } I've indicated 10 locations where I've seen validation checks placed in code. My question is what checks you would, if any, be performing at each given the following business rules (along with standard checks for length, range, format, type, etc): The amount of the cash drop must be greater than zero. The cash drop must have a valid Driver. The current user must be authorized to add cash drops (current user is not the driver). Please share your thoughts, how you have or would approach this scenario and the reasons for your choices.

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• Why RenderTarget2D overwrites other objects when trying to put some text in a model?

  - by cad

  I am trying to draw an object composited by two cubes (A & B) (one on top of the other, but for now I have them a little bit more open). I am able to do it and this is the result. (Cube A is the blue and Cube B is the one with brown text that comes from a png texture) But I want to have any text as parameter in the cube B. I have tried what @alecnash suggested in his question, but for some reason when I try to draw cube B, cube A dissapears and everything turns purple. This is my draw code: public void Draw(GraphicsDevice graphicsDevice, SpriteBatch spriteBatch, Matrix viewMatrix, Matrix projectionMatrix) { graphicsDevice.BlendState = BlendState.Opaque; graphicsDevice.DepthStencilState = DepthStencilState.Default; graphicsDevice.RasterizerState = RasterizerState.CullCounterClockwise; graphicsDevice.SamplerStates[0] = SamplerState.LinearClamp; // CUBE A basicEffect.View = viewMatrix; basicEffect.Projection = projectionMatrix; basicEffect.World = Matrix.CreateTranslation(ModelPosition); basicEffect.VertexColorEnabled = true; foreach (EffectPass pass in basicEffect.CurrentTechnique.Passes) { pass.Apply(); drawCUBE_TOP(graphicsDevice); drawCUBE_Floor(graphicsDevice); DrawFullSquareStripesFront(graphicsDevice, _numStrips, Color.Red, Color.Blue, _levelPercentage); DrawFullSquareStripesLeft(graphicsDevice, _numStrips, Color.Red, Color.Blue, _levelPercentage); DrawFullSquareStripesRight(graphicsDevice, _numStrips, Color.Red, Color.Blue, _levelPercentage); DrawFullSquareStripesBack(graphicsDevice, _numStrips, Color.Red, Color.Blue, _levelPercentage); } // CUBE B // Set the World matrix which defines the position of the cube texturedCubeEffect.World = Matrix.CreateTranslation(ModelPosition); // Set the View matrix which defines the camera and what it's looking at texturedCubeEffect.View = viewMatrix; // Set the Projection matrix which defines how we see the scene (Field of view) texturedCubeEffect.Projection = projectionMatrix; // Enable textures on the Cube Effect. this is necessary to texture the model texturedCubeEffect.TextureEnabled = true; Texture2D a = SpriteFontTextToTexture(graphicsDevice, spriteBatch, arialFont, "TEST ", Color.Black, Color.GhostWhite); texturedCubeEffect.Texture = a; //texturedCubeEffect.Texture = cubeTexture; // Enable some pretty lights texturedCubeEffect.EnableDefaultLighting(); // apply the effect and render the cube foreach (EffectPass pass in texturedCubeEffect.CurrentTechnique.Passes) { pass.Apply(); cubeToDraw.RenderToDevice(graphicsDevice); } } private Texture2D SpriteFontTextToTexture(GraphicsDevice graphicsDevice, SpriteBatch spriteBatch, SpriteFont font, string text, Color backgroundColor, Color textColor) { Vector2 Size = font.MeasureString(text); RenderTarget2D renderTarget = new RenderTarget2D(graphicsDevice, (int)Size.X, (int)Size.Y); graphicsDevice.SetRenderTarget(renderTarget); graphicsDevice.Clear(Color.Transparent); spriteBatch.Begin(); //have to redo the ColorTexture //spriteBatch.Draw(ColorTexture.Create(graphicsDevice, 1024, 1024, backgroundColor), Vector2.Zero, Color.White); spriteBatch.DrawString(font, text, Vector2.Zero, textColor); spriteBatch.End(); graphicsDevice.SetRenderTarget(null); return renderTarget; } The way I generate texture with dynamic text is: Texture2D a = SpriteFontTextToTexture(graphicsDevice, spriteBatch, arialFont, "TEST ", Color.Black, Color.GhostWhite); After commenting several parts to see what caused the problem, it seems to be located in this line graphicsDevice.SetRenderTarget(renderTarget);

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• Come up with a real-world problem in which only the best solution will do (a problem from Introduction to algorithms) [closed]

  - by Mike

  EDITED (I realized that the question certainly needs a context) The problem 1.1-5 in the book of Thomas Cormen et al Introduction to algorithms is: "Come up with a real-world problem in which only the best solution will do. Then come up with one in which a solution that is “approximately” the best is good enough." I'm interested in its first statement. And (from my understanding) it is asked to name a real-world problem where only the exact solution will work as opposed to a real-world problem where good-enough solution will be ok. So what is the difference between the exact and good enough solution. Consider some physics problem for example the simulation of the fulid flow in the permeable medium. To make this simulation happen some simplyfing assumptions have to be made when deriving a mathematical model. Otherwise the model becomes at least complex and unsolvable. Virtually any particle in the universe has its influence on the fluid flow. But not all particles are equal. Those that form the permeable medium are much more influental than the ones located light years away. Then when the mathematical model needs to be solved an exact solution can rarely be found unless the mathematical model is simple enough (wich probably means the model isn't close to reality). We take an approximate numerical method and after hours of coding and days of verification come up with the program or algorithm which is a solution. And if the model and an algorithm give results close to a real problem by some degree that is good enough soultion. Its worth noting the difference between exact solution algorithm and exact computation result. When considering real-world problems and real-world computation machines I believe all physical problems solutions where any calculations are taken can not be exact because universal physical constants are represented approximately in the computer. Any numbers are represented with the limited precision, at least limited by amount of memory available to computing machine. I can imagine plenty of problems where good-enough, good to some degree solution will work, like train scheduling, automated trading, satellite orbit calculation, health care expert systems. In that cases exact solutions can't be derived due to constraints on computation time, limitations in computer memory or due to the nature of problems. I googled this question and like what this guy suggests: there're kinds of mathematical problems that need exact solutions (little note here: because the question is taken from the book "Introduction to algorithms" the term "solution" means an algorithm or a program, which in this case gives exact answer on each input). But that's probably more of theoretical interest. So I would like to narrow down the question to: What are the real-world practical problems where only the best (exact) solution algorithm or program will do (but not the good-enough solution)? There are problems like breaking of cryptographic ciphers where only exact solution matters in practice and again in practice the process of deciphering without knowing a secret should take reasonable amount of time. Returning to the original question this is the problem where good-enough (fast-enough) solution will do there's no practical need in instant crack though it's desired. So the quality of "best" can be understood in any sense: exact, fastest, requiring least memory, having minimal possible network traffic etc. And still I want this question to be theoretical if possible. In a sense that there may be example of computer X that has limited resource R of amount Y where the best solution to problem P is the one that takes not more than available Y for inputs of size N*Y. But that's the problem of finding solution for P on computer X which is... well, good enough. My final thought that we live in a world where it is required from programming solutions to practical purposes to be good enough. In rare cases really very very good but still not the best ones. Isn't it? :) If it's not can you provide an example? Or can you name any such unsolved problem of practical interest?

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• F# Objects &ndash; Integration with the other .Net Languages &ndash; Part 2

  - by MarkPearl

  So in part one of my posting I covered the real basics of object creation. Today I will hopefully dig a little deeper… My expert F# book brings up an interesting point – properties in F# are just syntactic sugar for method calls. This makes sense… for instance assume I had the following object with the property exposed called Firstname. type Person(Firstname : string, Lastname : string) = member v.Firstname = Firstname I could extend the Firstname property with the following code and everything would be hunky dory… type Person(Firstname : string, Lastname : string) = member v.Firstname = Console.WriteLine("Side Effect") Firstname   All that this would do is each time I use the property Firstname, I would see the side effect printed to the screen saying “Side Effect”. Member methods have a very similar look & feel to properties, in fact the only difference really is that you declare that parameters are being passed in. type Person(Firstname : string, Lastname : string) = member v.FullName(middleName) = Firstname + " " + middleName + " " + Lastname   In the code above, FullName requires the parameter middleName, and if viewed from another project in C# would show as a method and not a property. Precomputation Optimizations Okay, so something that is obvious once you think of it but that poses an interesting side effect of mutable value holders is pre-computation of results. All it is, is a slight difference in code but can result in quite a huge saving in performance. Basically pre-computation means you would not need to compute a value every time a method is called – but could perform the computation at the creation of the object (I hope I have got it right). In a way I battle to differentiate this from lazy evaluation but I will show an example to explain the principle. Let me try and show an example to illustrate the principle… assume the following F# module namespace myNamespace open System module myMod = let Add val1 val2 = Console.WriteLine("Compute") val1 + val2 type MathPrecompute(val1 : int, val2 : int) = let precomputedsum = Add val1 val2 member v.Sum = precomputedsum type MathNormalCompute(val1 : int, val2 : int) = member v.Sum = Add val1 val2 Now assume you have a C# console app that makes use of the objects with code similar to the following… using System; using myNamespace; namespace CSharpTest { class Program { static void Main(string[] args) { Console.WriteLine("Constructing Objects"); var myObj1 = new myMod.MathNormalCompute(10, 11); var myObj2 = new myMod.MathPrecompute(10, 11); Console.WriteLine(""); Console.WriteLine("Normal Compute Sum..."); Console.WriteLine(myObj1.Sum); Console.WriteLine(myObj1.Sum); Console.WriteLine(myObj1.Sum); Console.WriteLine(""); Console.WriteLine("Pre Compute Sum..."); Console.WriteLine(myObj2.Sum); Console.WriteLine(myObj2.Sum); Console.WriteLine(myObj2.Sum); Console.ReadKey(); } } } The output when running the console application would be as follows…. You will notice with the normal compute object that the system would call the Add function every time the method was called. With the Precompute object it only called the compute method when the object was created. Subtle, but something that could lead to major performance benefits. So… this post has gone off in a slight tangent but still related to F# objects.

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• Play in NetBeans IDE (Part 2)

  - by Geertjan

  Peter Hilton was one of many nice people I met for the first time during the last few days constituting JAX London. He did a session today on the Play framework which, if I understand it correctly, is an HTML5 framework. It doesn't use web.xml, Java EE, etc. It uses Scala internally, as well as in its templating language.  Support for Play would, I guess, based on the little I know about it right now, consist of extending the HTML5 application project, which is new in NetBeans IDE 7.3. The workflow I imagine goes as follows. You'd create a new HTML5 application project, at which point you can choose a variety of frameworks and templates (Coffee Script, Angular, etc), which comes out of the box with the HTML5 support (i.e., Project Easel) in NetBeans IDE 7.3. Then, once the project is created, you'll right-click it and go to the Project Properties dialog, where you'll be able to enable Play support: At this stage, i.e., when you've checked the checkbox above and then clicked OK, all the necessary Play files will be added to your project, e.g., the routes file and the application.conf, for example. And then you have a Play application. Creating support in this way entails nothing more than creating a module that looks like this, i.e., with one Java class, where even the layer.xml file below is superfluous: All the code in the PlayEnablerPlanel.java that you see above is as follows: import java.awt.BorderLayout; import javax.swing.JCheckBox; import javax.swing.JComponent; import javax.swing.JPanel; import org.netbeans.spi.project.ui.support.ProjectCustomizer; import org.netbeans.spi.project.ui.support.ProjectCustomizer.Category; import org.openide.util.Lookup; public class PlayEnablerPanel implements ProjectCustomizer.CompositeCategoryProvider {     @ProjectCustomizer.CompositeCategoryProvider.Registration(             projectType = "org.netbeans.modules.web.clientproject",             position = 1000)     public static PlayEnablerPanel enablePlay() {         return new PlayEnablerPanel();     }     @Override     public Category createCategory(Lookup lkp) {         return ProjectCustomizer.Category.create("Play Framework", "Configure Play", null);     }     @Override     public JComponent createComponent(Category ctgr, Lookup lkp) {         JPanel playPanel = new JPanel(new BorderLayout());         playPanel.add(new JCheckBox("Enable Play"), BorderLayout.NORTH);         return playPanel;     } } Looking forward to having a beer with Peter soon (he lives not far away, in Rotterdam) to discuss this! Also read Part 1 of this series, which I wrote some time ago, and which has other ideas and considerations.

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• Too complex/too many objects?

  - by Mike Fairhurst

  I know that this will be a difficult question to answer without context, but hopefully there are at least some good guidelines to share on this. The questions are at the bottom if you want to skip the details. Most are about OOP in general. Begin context. I am a jr dev on a PHP application, and in general the devs I work with consider themselves to use many more OO concepts than most PHP devs. Still, in my research on clean code I have read about so many ways of using OO features to make code flexible, powerful, expressive, testable, etc. that is just plain not in use here. The current strongly OO API that I've proposed is being called too complex, even though it is trivial to implement. The problem I'm solving is that our permission checks are done via a message object (my API, they wanted to use arrays of constants) and the message object does not hold the validation object accountable for checking all provided data. Metaphorically, if your perm containing 'allowable' and 'rare but disallowed' is sent into a validator, the validator may not know to look for 'rare but disallowed', but approve 'allowable', which will actually approve the whole perm check. We have like 11 validators, too many to easily track at such minute detail. So I proposed an AtomicPermission class. To fix the previous example, the perm would instead contain two atomic permissions, one wrapping 'allowable' and the other wrapping 'rare but disallowed'. Where previously the validator would say 'the check is OK because it contains allowable,' now it would instead say '"allowable" is ok', at which point the check ends...and the check fails, because 'rare but disallowed' was not specifically okay-ed. The implementation is just 4 trivial objects, and rewriting a 10 line function into a 15 line function. abstract class PermissionAtom { public function allow(); // maybe deny() as well public function wasAllowed(); } class PermissionField extends PermissionAtom { public function getName(); public function getValue(); } class PermissionIdentifier extends PermissionAtom { public function getIdentifier(); } class PermissionAction extends PermissionAtom { public function getType(); } They say that this is 'not going to get us anything important' and it is 'too complex' and 'will be difficult for new developers to pick up.' I respectfully disagree, and there I end my context to begin the broader questions. So the question is about my OOP, are there any guidelines I should know: is this too complicated/too much OOP? Not that I expect to get more than 'it depends, I'd have to see if...' when is OO abstraction too much? when is OO abstraction too little? how can I determine when I am overthinking a problem vs fixing one? how can I determine when I am adding bad code to a bad project? how can I pitch these APIs? I feel the other devs would just rather say 'its too complicated' than ask 'can you explain it?' whenever I suggest a new class.

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• Is the Leptonica implementation of 'Modified Median Cut' not using the median at all?

  - by TheCodeJunkie

  I'm playing around a bit with image processing and decided to read up on how color quantization worked and after a bit of reading I found the Modified Median Cut Quantization algorithm. I've been reading the code of the C implementation in Leptonica library and came across something I thought was a bit odd. Now I want to stress that I am far from an expert in this area, not am I a math-head, so I am predicting that this all comes down to me not understanding all of it and not that the implementation of the algorithm is wrong at all. The algorithm states that the vbox should be split along the lagest axis and that it should be split using the following logic The largest axis is divided by locating the bin with the median pixel (by population), selecting the longer side, and dividing in the center of that side. We could have simply put the bin with the median pixel in the shorter side, but in the early stages of subdivision, this tends to put low density clusters (that are not considered in the subdivision) in the same vbox as part of a high density cluster that will outvote it in median vbox color, even with future median-based subdivisions. The algorithm used here is particularly important in early subdivisions, and 3is useful for giving visible but low population color clusters their own vbox. This has little effect on the subdivision of high density clusters, which ultimately will have roughly equal population in their vboxes. For the sake of the argument, let's assume that we have a vbox that we are in the process of splitting and that the red axis is the largest. In the Leptonica algorithm, on line 01297, the code appears to do the following Iterate over all the possible green and blue variations of the red color For each iteration it adds to the total number of pixels (population) it's found along the red axis For each red color it sum up the population of the current red and the previous ones, thus storing an accumulated value, for each red note: when I say 'red' I mean each point along the axis that is covered by the iteration, the actual color may not be red but contains a certain amount of red So for the sake of illustration, assume we have 9 "bins" along the red axis and that they have the following populations 4 8 20 16 1 9 12 8 8 After the iteration of all red bins, the partialsum array will contain the following count for the bins mentioned above 4 12 32 48 49 58 70 78 86 And total would have a value of 86 Once that's done it's time to perform the actual median cut and for the red axis this is performed on line 01346 It iterates over bins and check they accumulated sum. And here's the part that throws me of from the description of the algorithm. It looks for the first bin that has a value that is greater than total/2 Wouldn't total/2 mean that it is looking for a bin that has a value that is greater than the average value and not the median ? The median for the above bins would be 49 The use of 43 or 49 could potentially have a huge impact on how the boxes are split, even though the algorithm then proceeds by moving to the center of the larger side of where the matched value was.. Another thing that puzzles me a bit is that the paper specified that the bin with the median value should be located, but does not mention how to proceed if there are an even number of bins.. the median would be the result of (a+b)/2 and it's not guaranteed that any of the bins contains that population count. So this is what makes me thing that there are some approximations going on that are negligible because of how the split actually takes part at the center of the larger side of the selected bin. Sorry if it got a bit long winded, but I wanted to be as thoroughas I could because it's been driving me nuts for a couple of days now ;)

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• Asus X202e VivoBook, dual boot. How to get around UEFI and have Win8 & Ubuntu?

  - by Nukeface

  I've gotten my hands on an Asus Vivobook X202e. I like it, handy to use, small, etc etc. Oh, it's the i3 core version. For school I still need Windows * sigh * for the .NET development. (I know, possible in Ubuntu, this n that, but for ease atm wanting to keep it with Win8). So. How to install both on this little thing? I've found a way into the BIOS (before splash screen, mash F2. Works only after reboot, not cold boot). But the whole boot loading setup is different than from what I know, and I must've messed up something because it's been "Attempting Repairs", "Analyzing hard disk", and a bunch of other things for the past 15 minutes. (All I've done is selected "disabled" on secure boot, picky as ** Microsoft). Keeping the original Windows installation is of no concern. Found the product key already and have a clean install waiting. BTW, not trying to leech knowledge, even though first question and no answers. I'm more and more active on Stackoverflow. But, especially due to secure boot and windows 8, I'm going over to Ubuntu. Well, more and more anyway, I like my Windows based games as well ;) UPDATE Managed to do a clean install of Windows 8 Pro. After disabling Secure Boot, also had to disable fast boot, and enable Launch CSM, leaving the option which appeared (Launch PXE OpROM) disabled. Then I rebooted, with the USB Boot drive I created using the Windows 7 USB DVD Download Tool (scroll down for download link), provided by Microsoft. During the installation, I chose to install a clean version, therefor deleted the partitions containing current windows files. I left the Recovery partition (you never know...). Of course, the new Windows Installation dit not like this. Apparantly Windows cannot be installed on a GPT hard disk. Remember I hadn't changed the partition table, was still factory default! Minus a few partitions, granted. So deleted ALL partittions, did a format of the disk, created a new partition. Et voila, Windows installation started. FINALLY! WONDROUS After the installation, Windows still had background images located in C:/Users/ ME /AppData/Local/Microsoft/Themes/RoamedThemeFiles/DesktopBackground/ that I had in the previous installation. Before doing: format, delete partition, cascade partitions, create new partition of different size, format partition, install Windows. It managed to keep the images through all that. Anyone got an idea on that one? It also remembered the settings for the Windows Aero theme... UPDATED QUESTION: After all this you'd think I'd have the rest figured out. Wrong. Ubuntu 12.10, 64 bit installation can't read the partitioning of the hdd during the installation. Any ideas on how to fix this so the install for a dual-boot system can proceed? (Preferably without starting anew with Windows as well ;) )

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• Changing the Operating System with only Ubuntu installed

  - by Games Brainiac

  I really wanted to dive into the world of Open Source operating systems, so I downloaded the latest version of Ubuntu (13.10), and installed it on a clean(no operating system installed, absolutely nothing) Lenovo ThinkPad machine. After a few days, I wanted to try out a different Operating System (Elementary OS). I downloaded the ISO file, burned it to a USB, tested that the USB booted from a different computer (I have 2, one is the Lenovo, the other a HP). I was able to get the bootscreen, and everything worked like a charm after I set the BIOS to boot from USB Disk Drive instead of HD. After this, I went back to Lenovo, and tried to open up the boot menu, by pressing F12, so that I could load from a temporary device. To my surprise, nothing but the HD was listed. There was no Optical Drive, No USB Drive, absolutely nothing. So, I thought that these devices were probably disabled. So I went into my BIOS and checked to see what was the case. I saw that all my devices were enabled. USB and all the other devices such as network cable and the rest were all enabled. So, I thought this probably had something to do wit UEFI and Legacy Boot options. So, I made sure that both were enabled. This did not solve the problem either. Again, I got nothing but the option to boot from my Hard Disk. I thought the USB had to be at fault. I tried different ports, but to no avail. Next, I tried with a Live CD, which had Ubuntu on it. This failed too. I simply could not boot from anything other than my hard disk. Okay, so at this point, I was pretty desperate, so I installed Boot-Repair through: sudo add-apt-repository ppa:yannubuntu/boot-repair sudo apt-get update sudo apt-get install boot-repair What this did is lead me to GRUB. Ideally, its just a screen that gives me the option to load from Ubuntu or Advanced Settings. The Advanced settings had nothing but Ubuntu options in it. So, I kept on pressing ESC and that led me to the the grub console, and thats where I am right now with my Lenovo. I've also tried updating the BIOS, but Lenovo only has packages for Red Hat and Windows. So, a dead end there too. Right now, I need to know if there is any way that I can just delete everything from my Lenovo? I want to revert it back to its blank factory condition. How can I achieve this? I have tried to elaborate my problem as best I could. If there is any important information that I've missed out, please do not hesitate to leave a comment. I would have included some screen shots, but BIOS screen shots are a little hard to manage. However, I can provide a camera Image of the boot screen if needed (doing that as we speak).

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• Is there any kind of established architecture for browser based games?

  - by black_puppydog

  I am beginning the development of a broser based game in which players take certain actions at any point in time. Big parts of gameplay will be happening in real life and just have to be entered into the system. I believe a good kind of comparison might be a platform for managing fantasy football, although I have virtually no experience playing that, so please correct me if I am mistaken here. The point is that some events happen in the program (i.e. on the server, out of reach for the players) like pulling new results from some datasource, starting of a new round by a game master and such. Other events happen in real life (two players closing a deal on the transfer of some team member or whatnot - again: have never played fantasy football) and have to be entered into the system. The first part is pretty easy since the game masters will be "staff" and thus can be trusted to a certain degree to not mess with the system. But the second part bothers me quite a lot, especially since the actions may involve multiple steps and interactions with different players, like registering a deal with the system that then has to be approved by the other party or denied and passed on to a game master to decide. I would of course like to separate the game logic as far as possible from the presentation and basic form validation but am unsure how to do this in a clean fashion. Of course I could (and will) put some effort into making my own architectural decisions and prototype different ideas. But I am bound to make some stupid mistakes at some point, so I would like to avoid some of that by getting a little "book smart" beforehand. So the question is: Is there any kind of architectural works that I can read up on? Papers, blogs, maybe design documents or even source code? Writing this down this seems more like a business application with business rules, workflows and such... Any good entry points for that? EDIT: After reading the first answers I am under the impression of having made a mistake when including the "MMO" part into the title. The game will not be all fancy (i.e. 3D or such) on the client side and the logic will completely exist on the server. That is, apart from basic form validation for the user which will also be mirrored on the server side. So the target toolset will be HTML5, JavaScript, probably JQuery(UI). My question is more related to the software architecture/design of a system that enforces certain rules. Separation of ruleset and presentation One problem I am having is that I want to separate the game rules from the presentation. The first step would be to make an own module for the game "engine" that only exposes an interface that allows all actions to be taken in a clean way. If an action fails with regard to some pre/post condition, the engine throws an exception which is then presented to the user like "you cannot sell something you do not own" or "after that you would end up in a situation which is not a valid game state." The problem here is that I would like to be able to not even present invalid action in the first place or grey out the corresponding UI elements. Changing and tweaking the ruleset Another big thing is the ruleset. It will probably evolve over time and most definitely must be tweaked. What's more, it should be possible (to a certain extent) to build a ruleset that fits a specific game round, i.e. choosing different kinds of behaviours in different aspects of the game. This would do something like "we play it with extension A today but we throw out extension B." For me, this screams "Architectural/Design pattern" but I have no idea on who might have published on something like this, not even what to google for.

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• Is changing my job now a wise decision? [closed]

  - by FlaminPhoenix

  First a little background about myself. I am a javascript programmer with 3.8 years of experience. I joined my current company a year and 3 months ago, and I was recruited as a javascript programmer. I was under the impression I was a programmer in a programming team but this was not the case. No one else except me and my manager knows anything about programming in my team. The other two teammates, copy paste stuff from websites into excel sheets. I was told I was being recruited for a new project, and it was true. The only problem was that the server side language they were using was PHP. They were using a popular library with PHP, and I had never worked with PHP before. Nevertheless, I learnt it well enough to get things working, and received high praise from my boss's boss on whichever project I worked on. Words like "wow" , "This looks great, the clients gonna be impressed with this." were sprinkled every now and then on reviewing my work. They even managed to sell my work to a couple of clients and as I understand, both of my projects are going to fetch them a pretty buck. The problem: I was asked to move into a project which my manager was handling. I asked them for training on the project which never came, and sure enough I couldnt complete my first task on the new project without shortcomings. I told my manager there were things I didnt know how to get done in the new project due to lack of training. His project had 0 documentation. I was told he would "take care" of everything relating to those shortcomings. In the meantime, I was asked to switch to another project. My manager made the necessary changes and later told me that the build had "broken" on the production server and that I needed to "test" my changes before saying things were done. I never deployed it on the production server. He did. I never saw / had the opportunity to see the final build before it went to production. He called me for a separate meeting and started pointing fingers at me, but I took full responsibility even if I didnt have to. He later on got on a call with his boss, in my presence, and gave him the impression that it was all my fault. I did not confront him about this so far. I have worked late / done overtime without them asking a lot, but last week, I just got home from work, and I got calls asking me to solve an issue which till then they had kept quiet about even though they were informed about it. I asked my manager why I hadnt been tasked with this when I was in office. He started telling me which statements to put where, as if to mock me, and that this "is hardly an overtime issue" and this pissed me off. Also, during the previous meeting, he was constantly talking highly about his work, at the same time trying to demean mine. In the meantime, I have attended an interview with another MNC, and the interviewers there were fully respectful of my decision to leave my current company. Its a software company, so I can expect my colleagues to know a lot more than me. Im told I can expect their offer anytime this week. My questions: Is my anger towards my manager justified? While leaving, do I tell him that its because of his actions that Im leaving? Do I erupt in anger and tell him that he shouldnt have put the blame on me since he was the one doing the deployment? This is going to be my second resignation to this company. The first time I wanted to resign, I was asked to stay back and my manager promised a lot of changes, a couple of which were made. How do I keep myself from getting into such situations with my employers in the future?

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• I thought the new AUTO_SAMPLE_SIZE in Oracle Database 11g looked at all the rows in a table so why do I see a very small sample size on some tables?

  - by Maria Colgan

  I recently got asked this question and thought it was worth a quick blog post to explain in a little more detail what is going on with the new AUTO_SAMPLE_SIZE in Oracle Database 11g and what you should expect to see in the dictionary views. Let’s take the SH.CUSTOMERS table as an example.  There are 55,500 rows in the SH.CUSTOMERS tables. If we gather statistics on the SH.CUSTOMERS using the new AUTO_SAMPLE_SIZE but without collecting histogram we can check what sample size was used by looking in the USER_TABLES and USER_TAB_COL_STATISTICS dictionary views. The sample sized shown in the USER_TABLES is 55,500 rows or the entire table as expected. In USER_TAB_COL_STATISTICS most columns show 55,500 rows as the sample size except for four columns (CUST_SRC_ID, CUST_EFF_TO, CUST_MARTIAL_STATUS, CUST_INCOME_LEVEL ). The CUST_SRC_ID and CUST_EFF_TO columns have no sample size listed because there are only NULL values in these columns and the statistics gathering procedure skips NULL values. The CUST_MARTIAL_STATUS (38,072) and the CUST_INCOME_LEVEL (55,459) columns show less than 55,500 rows as their sample size because of the presence of NULL values in these columns. In the SH.CUSTOMERS table 17,428 rows have a NULL as the value for CUST_MARTIAL_STATUS column (17428+38072 = 55500), while 41 rows have a NULL values for the CUST_INCOME_LEVEL column (41+55459 = 55500). So we can confirm that the new AUTO_SAMPLE_SIZE algorithm will use all non-NULL values when gathering basic table and column level statistics. Now we have clear understanding of what sample size to expect lets include histogram creation as part of the statistics gathering. Again we can look in the USER_TABLES and USER_TAB_COL_STATISTICS dictionary views to find the sample size used. The sample size seen in USER_TABLES is 55,500 rows but if we look at the column statistics we see that it is same as in previous case except  for columns  CUST_POSTAL_CODE and  CUST_CITY_ID. You will also notice that these columns now have histograms created on them. The sample size shown for these columns is not the sample size used to gather the basic column statistics. AUTO_SAMPLE_SIZE still uses all the rows in the table - the NULL rows to gather the basic column statistics (55,500 rows in this case). The size shown is the sample size used to create the histogram on the column. When we create a histogram we try to build it on a sample that has approximately 5,500 non-null values for the column.  Typically all of the histograms required for a table are built from the same sample. In our example the histograms created on CUST_POSTAL_CODE and the CUST_CITY_ID were built on a single sample of ~5,500 (5,450 rows) as these columns contained only non-null values. However, if one or more of the columns that requires a histogram has null values then the sample size maybe increased in order to achieve a sample of 5,500 non-null values for those columns. n addition, if the difference between the number of nulls in the columns varies greatly, we may create multiple samples, one for the columns that have a low number of null values and one for the columns with a high number of null values.  This scheme enables us to get close to 5,500 non-null values for each column. +Maria Colgan

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• Installing WindowsAuthentication breaks authentication / web.config?

  - by Ian Quigley

  I have a clean Windows 2008 R2 box (on a VM) and have installed IIS 7.5 with default options. I then copied a website to it (from Windows 7, IIS 7) and after a little tweaking the website is working fine. The website is currently using and working with Anonymous Authentication. I have gone back to the Windows Components/Sever Manager, Roles - Security and ticked and installed Windows Authentication. When I check my server in IIS (top level above sites) - Authentication, I see Anonymous Authentication (enabled) ASP.NET Impersonation (disabled) Forms Authentication (disbaled) Windows Authentication (enabled) When I check my default website - Authentication, I see as above but "Retrieving status" and an error dialog saying There was an error while performing this operation. Details: Filename c:\inetpub\wwwroot\screwturnwiki\web.config Line number: 96 Error: This configuration section cannot be used in this path. This happens when the section is being locked at the parent level. Locking is either by default (overriderModeDefault="Deny"), or set explicity by a location tag with overrideMode="Deny" or the legacy allowOverride="False". I have tried hand editing the web.config with no success. (How to use locking in IIS7 Configuration) UN-installing Windows Authentication happily returns my site to working with Anonymous Authentication, and allows me to enable/disable these three options. FYI. I am using ScrewTurnWiki with the Active Directory plug in. It all works fine under Windows 7 IIS 7 locally (has been for months) Web.Config <system.webServer> (edit) <handlers> ( deleted removes/adds ) </handlers> <security> <authentication> 96: <windowsAuthentication enabled="true" useKernelMode="true"> <extendedProtection tokenChecking="Allow" /> <providers> <clear /> <add value="NTLM" /> <add value="Negotiate" /> </providers> </windowsAuthentication> </authentication> </security>

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• RDS installation failure on 2012 R2 Server Core VM in Hyper-V Server

  - by Giles

  I'm currently installing a test-bed for my firms Infrastructure replacement. 10 or so Windows/Linux servers will be replaced by 2 physical servers running Hyper-V server. All services (DC, RDS, SQL) will be on Windows 2012 R2 Server Core VMs, Exchange on Server 2012 R2 GUI, and the rest are things like Elastix, MailArchiver etc, which aren't part of the equation thus far. I have installed Hyper-V server on a test box, and sucessfully got two virtual DC's running, SQL 2014 running, and 8.1 which I use for the RSAT tools. When trying to install RDS (The old fashioned kind, not the newer VDI(?) style), I get a failed installation due to the server not being able to reboot. A couple of articles have said not to do it locally, so I've moved on. Sitting at the Powershell prompt on the Domain Controller or SQL server (Both Server Core), I run the following commands: Import-Module RemoteDesktop New-SessionDeployment -ConnectionBroker "AlstersTS.Alsters.local" -SessionHost "AlstersTS.Alsters.local" The installation begins, carries on for 2 or 3 minutes, then I receive the following error message: New-SessionDeployment : Validation failed for the "RD Connection Broker" parameter. AlstersTS.Alsters.local Unable to connect to the server by using WindowsPowerShell remoting. Verify that you can connect to the server. At line:1 char:1 + NewSessionDeployment -ConnectionBroker "AlstersTS.Alsters.local" -SessionHost " ... + + CategoryInfo : NotSpecified: (:) [Write-Error], WriteErrorException + FullyQualifiedErrorID : Microsoft.PowerShell.Commands.WriteErrorException,New-SessionDeployment So far, I have: Triple, triple checked syntax. Tried various other commands, and a script to accomplish the same task. Checked DNS is functioning as it should. Checked to the best of my knowledge that AD is working as it should. Checked that the Network Service has the needed permissions. Created another VM and placed the two roles on different servers. Deleted all VMs, started again with a new domain name (Lather, rinse, repeat) Performed the whole installation on a second physical box running Hyper-V Server Pleaded with it Interestingly, if I perform the installation via a GUI installation, the thing just works! Now I know I could convert this to a Server Core role after installation, but this wouldn't teach me what was wrong in the first instance. I've probably got 10 pages through various Google searches, each page getting a little less relevant. The closest matches seem to have good information, but it doesn't seem to be the fix for my set-up. As a side note, I expected to be able to "tee" or "out-file" the error message into a text file, but couldn't get that to work either, so I've typed in the error message manually. Chaps, any suggestions, from the glaringly obvious, to the long-winded and complex? Thanks!

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• Migrate from MySQL to PostgreSQL on Linux (Kubuntu)

  - by Dave Jarvis

  Storyline Trying to migrate a database from MySQL to PostgreSQL. All the documentation I have read covers, in great detail, how to migrate the structure. I have found very little documentation on migrating the data. The schema has 13 tables (which have been migrated successfully) and 9 GB of data. MySQL version: 5.1.x PostgreSQL version: 8.4.x I want to use the R programming language to analyze the data using SQL select statements; PostgreSQL has PL/R, but MySQL has nothing (as far as I can tell). A long time ago in a galaxy far, far away... Create the database location (/var has insufficient space; also dislike having the PostgreSQL version number everywhere -- upgrading would break scripts!): sudo mkdir -p /home/postgres/main sudo cp -Rp /var/lib/postgresql/8.4/main /home/postgres sudo chown -R postgres.postgres /home/postgres sudo chmod -R 700 /home/postgres sudo usermod -d /home/postgres/ postgres All good to here. Next, restart the server and configure the database using these installation instructions: sudo apt-get install postgresql pgadmin3 sudo /etc/init.d/postgresql-8.4 stop sudo vi /etc/postgresql/8.4/main/postgresql.conf Change data_directory to /home/postgres/main sudo /etc/init.d/postgresql-8.4 start sudo -u postgres psql postgres \password postgres sudo -u postgres createdb climate pgadmin3 Use pgadmin3 to configure the database and create a schema. A New Hope The episode began in a remote shell known as bash, with both databases running, and the installation of a command with a most unusual logo: SQL Fairy. perl Makefile.PL sudo make install sudo apt-get install perl-doc (strangely, it is not called perldoc) perldoc SQL::Translator::Manual Extract a PostgreSQL-friendly DDL and all the MySQL data: sqlt -f DBI --dsn dbi:mysql:climate --db-user user --db-password password -t PostgreSQL > climate-pg-ddl.sql mysqldump --skip-add-locks --complete-insert --no-create-db --no-create-info --quick --result-file="climate-my.sql" --databases climate --skip-comments -u root -p The Database Strikes Back Recreate the structure in PostgreSQL as follows: pgadmin3 (switch to it) Click the Execute arbitrary SQL queries icon Open climate-pg-ddl.sql Search for TABLE " replace with TABLE climate." (insert the schema name climate) Search for on " replace with on climate." (insert the schema name climate) Press F5 to execute This results in: Query returned successfully with no result in 122 ms. Replies of the Jedi At this point I am stumped. Where do I go from here (what are the steps) to convert climate-my.sql to climate-pg.sql so that they can be executed against PostgreSQL? How to I make sure the indexes are copied over correctly (to maintain referential integrity; I don't have constraints at the moment to ease the transition)? How do I ensure that adding new rows in PostgreSQL will start enumerating from the index of the last row inserted (and not conflict with an existing primary key from the sequence)? Resources A fair bit of information was needed to get this far: https://help.ubuntu.com/community/PostgreSQL http://articles.sitepoint.com/article/site-mysql-postgresql-1 http://wiki.postgresql.org/wiki/Converting_from_other_Databases_to_PostgreSQL#MySQL http://pgfoundry.org/frs/shownotes.php?release_id=810 http://sqlfairy.sourceforge.net/ Thank you!

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• PXE-E32 TFTP Open Timeout While Attempting to PXE Boot from Windows Deployment Services

  - by bschafer

  I'm running Windows Deployment Services on Windows Server 2008 R2 on top of an ESX 4.0 box. This is the only function of this VM instance, although it had previously functioned as an AD Domain Controller. My DHCP server is running on our primary Domain Controller, which is also Server 2008 R2, but running on metal. Everything was working perfectly until we recently had our backup generator fail during a power outage, causing all of our servers and networking equipment to lose power for a period of time. When we brought all of our equipment back up, everything was working as expected except for WDS. Our network is split up into several different vlans. Now, depending on which vlan the client computer is on, it's behaving differently when attempting to PXE boot into WDS. Our servers are located on the 10.55.x.x vlan, which, due to the nature of it, has no DHCP server active in it. The first computer we plugged in happened to be in the 10.99.x.x vlan, which is supposed to be reserved for network management devices (i.e. switches), but we've been using it occasionally otherwise. That computer gave us PXE-E11 ARP Timeout errors. When we moved to a different computer on the 10.19.x.x vlan (for general purpose use), it finally gets an IP from DHCP, but it presents us with a very stumping PXE-E32 TFTP Open Timeout error. Before the power outage, it didn't matter which vlan a device was on; it would PXE boot and image just fine. I've made no changes to anything server-side. Everything is configured exactly the same way it was on my WDS and DHCP servers as before the power outage. I've tried several different computers, including different models. All of this, combined with the quirky behavior depending on the vlan, makes me think something went wrong in one or more of our switches, probably because of the power outage. Unfortunately, I'm no network guy, and I know very little about how to configure our switches properly. Is this an issue with switches, etc? If so, how can I fix it? Is there some magical option I'm not aware of? Does anybody out there have any hunches? I've pretty much exhausted my ideas. Our main switch is an HP Procurve 5406. We also have 3x HP Procurve 4208 switches. The ESX Server is an HP ProLiant DL380 G6. The WDS VM is currently using the VMXNET3 network adaptor, but we've also tried the E1000 adaptor.

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• Best available technology for layered disk cache in linux

  - by SpliFF

  I've just bought a 6-core Phenom with 16G of RAM. I use it primarily for compiling and video encoding (and occassional web/db). I'm finding all activities get disk-bound and I just can't keep all 6 cores fed. I'm buying an SSD raid to sit between the HDD and tmpfs. I want to setup a "layered" filesystem where reads are cached on tmpfs but writes safely go through to the SSD. I want files (or blocks) that haven't been read lately on the SSD to then be written back to a HDD using a compressed FS or block layer. So basically reads: - Check tmpfs - Check SSD - Check HD And writes: - Straight to SSD (for safety), then tmpfs (for speed) And periodically, or when space gets low: - Move least frequently accessed files down one layer. I've seen a few projects of interest. CacheFS, cachefsd, bcache seem pretty close but I'm having trouble determining which are practical. bcache seems a little risky (early adoption), cachefs seems tied to specific network filesystems. There are "union" projects unionfs and aufs that let you mount filesystems over each other (USB device over a DVD usually) but both are distributed as a patch and I get the impression this sort of "transparent" mounting was going to become a kernel feature rather than a FS. I know the kernel has a built-in disk cache but it doesn't seem to work well with compiling. I see a 20x speed improvement when I move my source files to tmpfs. I think it's because the standard buffers are dedicated to a specific process and compiling creates and destroys thousands of processes during a build (just guessing there). It looks like I really want those files precached. I've read tmpfs can use virtual memory. In that case is it practical to create a giant tmpfs with swap on the SSD? I don't need to boot off the resulting layered filesystem. I can load grub, kernel and initrd from elsewhere if needed. So that's the background. The question has several components I guess: Recommended FS and/or block layer for the SSD and compressed HDD. Recommended mkfs parameters (block size, options etc...) Recommended cache/mount technology to bind the layers transparently Required mount parameters Required kernel options / patches, etc..

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• IIS 7.5 FTPS external access - 534 Policy requires SSL

  - by markmnl

  I have setup a FTP site that requires SSL but when I try connect to it externally I get the error: 220 Microsoft FTP Service 534 Policy requires SSL. I know - I set it so! Why doesnt it fetch the SSL cert from the site and allow me to logon?! (Incidentally beware of all the tutorials that Allow but do not Require SSL - while that will solve the problem it will be because SSL is not being used!). I suspect it may be I need a client that supports FTPS (FTP over SSL) and Windows explorer just uses IE which does not. But trying FileZilla and WinSCP I get a little further but then it hangs on TLS/SSL negotiation expecting a response from the server.... UPDATE: I have tried (from: http://learn.iis.net/page.aspx/309/configuring-ftp-firewall-settings/): Configure the Passive Port Range for the FTP Service. Configure the external IPv4 Address for a Specific FTP Site. Configure the firewall to allow the FTP service to listen on all ports that it opens. Disabling stateful FTP filtering so that Windows Firewall will not block FTP traffic. And still I get (in FileZilla trying both Active and Passive): Status: Connecting to 203.x.x.x:21... Status: Connection established, waiting for welcome message... Response: 220 Microsoft FTP Service Command: AUTH TLS Response: 234 AUTH command ok. Expecting TLS Negotiation. Status: Initializing TLS... Error: Connection timed out Error: Could not connect to server The Windows firewall logs unhelpfully have nothing to say.. UPDATE2: Turning the firewall off does not resolve the problem. I cannot believe how difficult it is to get something so simple to work and even once following the documentation it does not work. UPDATE3: Running FileZilla locally connecting through the loopback works in Active mode, in Passive mode I get up to: Command: LIST Response: 150 Opening BINARY mode data connection. Error: GnuTLS error -53: Error in the push function. Turning the firewall off at both ends I can still not connect the client and get the same error as above.

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• Windows 7 is shutting down unexpectedly, according to the logs.

  - by dlamblin

  Here's a message from my eventvwr EventLog (Windows Logs System): The previous system shutdown at 11:51:15 AM on ?7/?29/?2009 was unexpected. This is funny because I was wondering why the system shut down while I was playing Civilizations IV full screen. Now I know. It was unexpected. Has anyone encountered and resolved this? A little background: I am running Windows 7 RC inside VMWare Fusion 2 (just updated a few months back) on a MacBook (Bitterly not Pro) aluminum-body. Windows 7 occasionally will shut down. This isn't a quick turn-off, it's a shutdown where all the programs are exited, the system waits until they quit (and Civ4 doesn't prompt me to save), it even installed Windows Updates before restarting. And yes it is restarting right after the shutdown. Because I run a game in full screen mode I do not notice any dialog with a countdown timer or anything like that that might be a warning. As I have iStat on my dashboard widgets I can see about 8 temperature monitors. I have seen the CPU get up to 74C before, but during the shutdown, though it seemed hot to the touch (always is), it read 61C for the CPU, 60C for heatsink A, 50C for heatsink B and in the 30s-40s for the enclosure and harddrives. As I type this now, the temps are actually higher, so I don't think the temperature caused it. I have at least six such events dating first from 5/17 which was a week after installing Windows 7. I did find one information level warning from USER32 in the system log that says: The process C:\Windows\system32\svchost.exe (DLAMBLIN-WIN7) has initiated the restart of computer DLAMBLIN-WIN7 on behalf of user NT AUTHORITY\SYSTEM for the following reason: Operating System: Recovery (Planned) Reason Code: 0x80020002 Shutdown Type: restart Comment: And another 15 minutes before that from Windows Update: Restart Required: To complete the installation of the following updates, the computer will be restarted within 15 minutes: - Cumulative Security Update for Internet Explorer 8 for Windows 7 Release Candidate for x64-based Systems (KB972260) Which I think kind of explains it. Though I don't know why restarting after an update would create an error event of "shutdown was unexpected", isn't that pretty odd? Now, how do I set it to never restart after an update unless I click something. Application of solution: As fretje reminded me, there's a couple of configurable settings for this, in windows 7 they're much in the same place as in Windows 2000 SP3 and XP SP1. Running gpedit.msc pops up a window that looks like: Windows 7 has changed the order and added a couple of newer options I've italicized: Do not display 'Install Updates and Shut Down' in Shut Down Windows dialog box Do not adjust default option to 'Install Updates and Shut Down' in Shut Down Windows dialog box Enabling Windows Power Management to automatically wake up the system to install scheduled updates Configure Automatic Updates Specify intranet Microsoft update service location Automatic Updates detection frequency Allow non-administrators to receive update notifications Turn on Software Notifications Allow Automatic Updates immediate installation Turn on recommended updates via Automatic Updates No auto-restart with logged-on users for scheduled Automatic Updates Re-prompt for restart with scheduled installations. Delay Restart for scheduled installations Reschedule Automatic Updates schedule

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• Getting macro keys from a razer blackwidow to work on linux

  - by Journeyman Geek

  I picked up a razer blackwidow ultimate that has additional keys meant for macros that are set using a tool that's installed on windows. I'm assuming that these arn't some fancypants joojoo keys and should emit scancodes like any other keys. Firstly is there a standard way to check these scancodes in linux? Secondly how do i set these keys to do things in command line and x based linux setups? My current linux install is xubuntu 10.10, but i'll be switching to kubuntu once i have a few things fixed up. Ideally the answer should be generic and system-wide Things i have tried so far: showkeys from the built in kbd package (in a seperate vt) - macro keys not detected xev - macro keys not detected lsusb and evdev output this ahk script's output suggests the M keys are not outputting standard scancodes Things i need to try snoopy pro + reverse engineering (oh dear) Wireshark - preliminary futzing around seems to indicate no scancodes emitted when what i seem to think is the keyboard is monitored and keys pressed. Might indicate additional keys are a seperate device or need to be initialised somehow. Need to cross reference that with lsusb output from linux, in 3 scenarios - standalone, passed through to a windows VM without the drivers installed, and the same with. LSUSB only detects one device on a standalone linux install It might be useful to check if the mice use the same razer synapse driver , since that means some variation of razercfg might work (not detected. only seems to work for mice) Things i have Have worked out: In a windows system with the driver, the keyboard is seen as a keyboard and a pointing device. And said pointing device uses, in addition to your bog standard mouse drivers.. a driver for something called a razer synapse. Mouse driver seen in linux under evdev and lsusb as well Single Device under OS X apparently, though i have yet to try lsusb equivilent on that Keyboard goes into pulsing backlight mode in OS X upon initialisation with the driver. This should probably indicate that there's some initialisation sequence sent to the keyboard on activation. They are, in fact, fancypants joojoo keys. Extending this question a little I have access to a windows system so if i need to use any tools on that to help answer the question, its fine. I can also try it on systems with and without the config utility. The expected end result is still to make those keys usable on linux however. I also realise this is a very specific family of hardware. I would be willing to test anything that makes sense on a linux system if i have detailed instructions - this should open up the question to people who have linux skills, but no access to this keyboard The minimum end result i require I need these keys detected, and usable in any fashion on any of the current graphical mainstream ubuntu varients

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• swapping or thrashing with vast amounts of unmapped pagecache

  - by Marco

  EDIT: I noticed that this is more appropriate for superuser.com, I apologize. I don't know how to delete this question. I'm using kubuntu jaunty (i386 32bit), kernel 2.6.28-13-generic. I've 4Gb of RAM, of which only 3317Mb are seen by the system (I guess because of the 32bit system). I'm seeing that the pagecache utilization is continually growing, up to the point that the system is unusable (after a few days). This happens also when I don't do anything (all user applications closed and the bare minimum of services enabled). If enabled, the system starts to use swap space (using it all in the end). Even if swap is disabled, disk activity becomes continuous, with the system unresponsive. For example, right now the system is working (albeit a tad slow), with only firefox and wing ide running, and I have 2Gb cached with only 45Mb mapped: $ free total used free shared buffers cached Mem: 3346388 3247328 99060 0 8416 2117980 -/+ buffers/cache: 1120932 2225456 Swap: 2144668 519448 1625220 $ cat /proc/meminfo MemTotal: 3346388 kB MemFree: 97128 kB Buffers: 7872 kB Cached: 2120224 kB SwapCached: 413860 kB Active: 2304596 kB Inactive: 865984 kB Active(anon): 2279168 kB Inactive(anon): 830236 kB Active(file): 25428 kB Inactive(file): 35748 kB Unevictable: 32 kB Mlocked: 32 kB HighTotal: 2492940 kB HighFree: 5456 kB LowTotal: 853448 kB LowFree: 91672 kB SwapTotal: 2144668 kB SwapFree: 1625244 kB Dirty: 84 kB Writeback: 0 kB AnonPages: 629304 kB Mapped: 45768 kB Slab: 45600 kB SReclaimable: 21756 kB SUnreclaim: 23844 kB PageTables: 4468 kB NFS_Unstable: 0 kB Bounce: 0 kB WritebackTmp: 0 kB CommitLimit: 3817860 kB Committed_AS: 3735020 kB VmallocTotal: 122880 kB VmallocUsed: 9352 kB VmallocChunk: 66600 kB HugePages_Total: 0 HugePages_Free: 0 HugePages_Rsvd: 0 HugePages_Surp: 0 Hugepagesize: 4096 kB DirectMap4k: 16376 kB DirectMap4M: 888832 kB If I try to drop the caches, little happes: # sync ; echo 3 > /proc/sys/vm/drop_caches ; free total used free shared buffers cached Mem: 3346388 3220580 125808 0 3020 2100600 -/+ buffers/cache: 1116960 2229428 Swap: 2144668 519356 1625312 Right now I've vm.swappiness = 5, but I've tried also with 0 and 1 (without noticeable differences). I've also tried vm.vfs_cache_pressure = 50 and 150 (again, no differences). As I said the pagecache eats all memory even with swapping turned off. What is happening? How to avoid this? TIA, Marco

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• svn 503 error when commit new files

  - by philipp

  I am struggling with a strange error when I try to commit my repository. I have a V-server with webmin installed on it. Via Webmin I installed an svn module, created repositories and everything worked fine until three days ago. Trying to commit brings the following error: Commit failed (details follow): Server sent unexpected return value (503 Service unavailable) in response to PROPFIND request for '/svn/rle/!svn/wrk/a1f963a7-0a33-fa48-bfde-183ea06ab958/RLE/.htaccess' Server sent unexpected return value (503 Service unavailable) in response to PROPFIND request for '/svn/rle/RLE/.htaccess' I google everywhere and found only very few solutions. One indicated that a wrong error document is set, another one dealt about the problem that filenames might cause this error and last but not least a wrong proxy configuration in the local svn config could be the reason. After trying all of the solutions suggested I could not reach anything. Only after a server reboot there was a little difference in the error-message, telling me that the server was not able to move a temp file, because the operation was permitted. So I also controlled the permissions of the svn directory, but also with no success. An svn update than restored the "normal" error from above and nothing changed since then. The only change I made on the server, I guess that this could be the reason why svn does not work anymore, was to install the php5_mysql module for apache via apt-get install php5_mysql. Atg the moment I have totally no idea where I could search. I don't know if the problem is on my server or in my repository and I would be glad to get any hint to solve this. Thanks in advance Greetings philipp error log: [Tue Oct 25 19:23:02 2011] [error] [client 217.50.254.18] Could not create activity /svn/rle/!svn/act/d8dd436f-d014-f047-8e87-01baac46a593. [500, #0] Tue Oct 25 19:23:02 2011] [error] [client 217.50.254.18] could not begin a transaction [500, #1] [Tue Oct 25 19:24:21 2011] [error] [client 217.50.254.18] Could not create activity /svn/rle/!svn/act/adac52c2-6f46-f540-b218-2f2ff03b51a4. [500, #0] http.conf: <Location /svn> DAV svn SVNParentPath /home/xxx/svn AuthType Basic AuthName xxx.de AuthUserFile /home/xxx/etc/svn.basic.passwd Require valid-user AuthzSVNAccessFile /home/xxx/etc/svn-access.conf Satisfy Any ErrorDocument 404 default RewriteEngine off </Location> The permissions for the repository directory are : rwxrwxrwx (0777). the directory /svn/rle/!svn/act/adac52c2-6f46-f540-b218-2f2ff03b51a4 does not exist on the server. I think this is part of the repository. So, I just just want to admit that i tried to reach the repository via Browser and i worked, I could see everything, so the error only occurs when I try to commit new files. I also created a second repository and tried to commit files in there, what gave me the same error.

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• Setting up a transparent SSL proxy

  - by badunk

  I've got a linux box set up with 2 network cards to inspect traffic going through port 80. One card is used to go out to the internet, the other one is hooked up to a networking switch. The point is to be able to inspect all HTTP and HTTPS traffic on devices hooked up to that switch for debugging purposes. I've written the following rules for iptables: nat -A PREROUTING -i eth1 -p tcp -m tcp --dport 80 -j DNAT --to-destination 192.168.2.1:1337 -A PREROUTING -i eth1 -p tcp -m tcp --dport 80 -j REDIRECT --to-ports 1337 -A POSTROUTING -s 192.168.2.0/24 -o eth0 -j MASQUERADE On 192.168.2.1:1337, I've got a transparent http proxy using Charles (http://www.charlesproxy.com/) for recording. Everything's fine for port 80, but when I add similar rules for port 443 (SSL) pointing to port 1337, I get an error about invalid message through Charles. I've used SSL proxying on the same computer before with Charles (http://www.charlesproxy.com/documentation/proxying/ssl-proxying/), but have been unsuccessful with doing it transparently for some reason. Some resources I've googled say its not possible - I'm willing to accept that as an answer if someone can explain why. As a note, I have full access to the described set up including all the clients hooked up to the subnet - so I can accept self-signed certs by Charles. The solution doesn't have to be Charles-specific since in theory, any transparent proxy will do. Thanks! Edit: After playing with it a little, I was able to get it working for a specific host. When I modify my iptables to the following (and open 1338 in charles for reverse proxy): nat -A PREROUTING -i eth1 -p tcp -m tcp --dport 80 -j DNAT --to-destination 192.168.2.1:1337 -A PREROUTING -i eth1 -p tcp -m tcp --dport 80 -j REDIRECT --to-ports 1337 -A PREROUTING -i eth1 -p tcp -m tcp --dport 443 -j DNAT --to-destination 192.168.2.1:1338 -A PREROUTING -i eth1 -p tcp -m tcp --dport 443 -j REDIRECT --to-ports 1338 -A POSTROUTING -s 192.168.2.0/24 -o eth0 -j MASQUERADE I am able to get a response, but with no destination host. In the reverse proxy, if I just specify that everything from 1338 goes to a specific host that I wanted to hit, it performs the hand shake properly and I can turn on SSL proxying to inspect the communication. The setup is less than ideal because I don't want to assume everything from 1338 goes to that host - any idea why the destination host is being stripped? Thanks again

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• Sendmail - Multiple Domains, One Box - Blocking One Or Two Domains

  - by TangoOversway

  I have a number of domains hosted at a web hosting service. They use sendmail to handle incoming email. I have six domains on this service (which we can call aaa.com, bbb.com and so on). Each email account has the same name and one email box. In other words, [email protected], [email protected], [email protected] and all the others go into one box, /var/spool/mail/tango, where my email program on my desktop picks it up. I have done very little work in sendmail. I haven't had to, and I've been warned it's a steep learning curve. But now I'm running into an issue. I was in a business situation where, for years, my email address was on the website for aaa.com. (We won't go into why this was necessary - it wasn't my preference and it's in the past.) Now I'm using [email protected] instead of [email protected]. I was getting about 1,000 or more pieces of spam a day, but SpamAssassin and my own email program caught about 75% of that. (Which still left stuff to delete.) Now, after checking, I see that 90% or more goes to [email protected], the one that was on the web for years. I'd like to deactivate [email protected] and possibly [email protected] and [email protected], but want to keep using [email protected]. Remember, email to tango at any of these domains will go into one email box. I've had people tell me that sendmail can be configured so I can deactivate [email protected] (and other domains) and still use [email protected] (and others, if I want to). In other words, I can configure sendmail to use this account on some domains and not others. One of the people who was teling me this was in tech support at the hosting service. But I wrote to tech support with a work order to do this and now I'm told it can't be done. I can modify config files myself on this account if needed, but I was hoping to just let them do it. (I love delegation -- it means I spend more time doing my stuff.) Is it possible to keep an email account active on one domain and not others with sendmail, when all domains are hosted on the same server? Is there a name for this process or setting? Any information would be helpful - either pointers to instructions so I can do it, or enough info so I can tell tech support, "This is where to look, and it can be done, so please pass my request on to someone who works with sendmail and knows how to do it." Is this something sendmail can do?

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