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  • Help with 2-part question on ASP.NET MVC and Custom Security Design

    - by JustAProgrammer
    I'm using ASP.NET MVC and I am trying to separate a lot of my logic. Eventually, this application will be pretty big. It's basically a SaaS app that I need to allow for different kinds of clients to access. I have a two part question; the first deals with my general design and the second deals with how to utilize in ASP.NET MVC Primarily, there will initially be an ASP.NET MVC "client" front-end and there will be a set of web-services for third parties to interact with (perhaps mobile, etc). I realize I could have the ASP.NET MVC app interact just through the Web Service but I think that is unnecessary overhead. So, I am creating an API that will essentially be a DLL that the Web App and the Web Services will utilize. The API consists of the main set of business logic and Data Transfer Objects, etc. (So, this includes methods like CreateCustomer, EditProduct, etc for example) Also, my permissions requirements are a little complicated. I can't really use a straight Roles system as I need to have some fine-grained permissions (but all permissions are positive rights). So, I don't think I can really use the ASP.NET Roles/Membership system or if I can it seems like I'd be doing more work than rolling my own. I've used Membership before and for this one I think I'd rather roll my own. Both the Web App and Web Services will need to keep security as a concern. So, my design is kind of like this: Each method in the API will need to verify the security of the caller In the Web App, each "page" ("action" in MVC speak) will also check the user's permissions (So, don't present the user with the "Add Customer" button if the user does not have that right but also whenever the API receives AddCustomer(), check the security too) I think the Web Service really needs the checking in the DLL because it may not always be used in some kind of pre-authenticated context (like using Session/Cookies in a Web App); also having the security checks in the API means I don't really HAVE TO check it in other places if I'm on a mobile (say iPhone) and don't want to do all kinds of checking on the client However, in the Web App I think there will be some duplication of work since the Web App checks the user's security before presenting the user with options, which is ok, but I was thinking of a way to avoid this duplication by allowing the Web App to tell the API not check the security; while the Web Service would always want security to be verified Is this a good method? If not, what's better? If so, what's a good way of implementing this. I was thinking of doing this: In the API, I would have two functions for each action: // Here, "Credential" objects are just something I made up public void AddCustomer(string customerName, Credential credential , bool checkSecurity) { if(checkSecurity) { if(Has_Rights_To_Add_Customer(credential)) // made up for clarity { AddCustomer(customerName); } else // throw an exception or somehow present an error } else AddCustomer(customerName); } public void AddCustomer(string customerName) { // actual logic to add the customer into the DB or whatever // Would it be good for this method to verify that the caller is the Web App // through some method? } So, is this a good design or should I do something differently? My next question is that clearly it doesn't seem like I can really use [Authorize ...] for determining if a user has the permissions to do something. In fact, one action might depend on a variety of permissions and the View might hide or show certain options depending on the permission. What's the best way to do this? Should I have some kind of PermissionSet object that the user carries around throughout the Web App in Session or whatever and the MVC Action method would check if that user can use that Action and then the View will have some ViewData or whatever where it checks the various permissions to do Hide/Show?

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  • [C#] Not enough memory or not enough handles?

    - by Nayan
    I am working on a large scale project where a custom (pretty good and robust) framework has been provided and we have to use that for showing up forms and views. There is abstract class StrategyEditor (derived from some class in framework) which is instantiated whenever a new StrategyForm is opened. StrategyForm (a customized window frame) contains StrategyEditor. StrategyEditor contains StrategyTab. StrategyTab contains StrategyCanvas. This is a small portion of the big classes to clarify that there are many objects that will be created if one StrategyForm object is allocated in memory at run-time. My component owns all these classes mentioned above except StrategyForm whose code is not in my control. Now, at run-time, user opens up many strategy objects (which trigger creation of new StrategyForm object.) After creating approx. 44 strategy objects, we see that the USER OBJECT HANDLES (I'll use UOH from here onwards) created by the application reaches to about 20k+, while in registry the default amount for handles is 10k. Read more about User Objects here. Testing on different machines made it clear that the number of strategy objects opened is different for message to pop-up - on one m/c if it is 44, then it can be 40 on another. When we see the message pop-up, it means that the application is going to respond slowly. It gets worse with few more objects and then creation of window frames and subsequent objects fail. We first thought that it was not-enough-memory issue. But then reading more about new in C# helped in understanding that an exception would be thrown if app ran out of memory. This is not a memory issue then, I feel (task manager also showed 1.5GB+ available memory.) M/C specs Core 2 Duo 2GHz+ 4GB RAM 80GB+ free disk space for page file Virtual Memory set: 4000 - 6000 My questions Q1. Does this look like a memory issue and I am wrong that it is not? Q2. Does this point to exhaustion of free UOHs (as I'm thinking) and which is resulting in failure of creation of window handles? Q3. How can we avoid loading up of an StrategyEditor object (beyond a threshold, keeping an eye on the current usage of UOHs)? (we already know how to fetch number of UOHs in use, so don't go there.) Keep in mind that the call to new StrategyForm() is outside the control of my component. Q4. I am bit confused - what are Handles to user objects exactly? Is MSDN talking about any object that we create or only some specific objects like window handles, cursor handles, icon handles? Q5. What exactly cause to use up a UOH? (almost same as Q4) I would be really thankful to anyone who can give me some knowledgeable answers. Thanks much! :)

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  • JQuery Help: slidetoggle() is not working properly

    - by John Smith
    Hi all, I'm trying to use JQuery Slidetoggle functionality, but not able to use properly. The problem I'm currently facing is my div is sliding down on the click of slide image icon, but after that suddenly data div (in which data is loading) disappears. Means sliding down is perfect but div (in which data is loading , here #divMain) is not visible after that. I want to achieve sliding effect in my code, like this has (Please see Website Design, Redesign Services slider): Here is my code: HTML: <div> <div class="jquery_inner_mid"> <div class="main_heading"> <a href="#"> <img src="features.jpg" alt="" title="" border="0" /></a> </div> <div class="plus_sign"> <img id="imgFeaturesEx" src="images/plus.jpg" alt="" title="" border="0" /> <img id="imgFeaturesCol" src="images/minus.jpg" alt="" title="" border="0" /></div> <div class="toggle_container"> <div id="divMain" > </div> </div> </div> <div class="jquery_inner_mid"> <div class="main_heading"> <img src="About.jpg" alt="" title="" /></div> <div class="plus_sign"> <img id="imgTechnoEx" src="images/plus.jpg" alt="" title="" border="0" /> <img id="imgTechnoCol" src="images/minus.jpg" alt="" title="" border="0" /></div> <div class="toggle_container"> <div id="divTechnossus" > </div> </div> </div> </div> JQuery: $(function() { $('#imgFeaturesCol').hide(); $('#imgTechnoCol').hide(); $('#imgFeaturesEx').click(function() { $.getJSON("/Visitor/GetFeatureInfo", null, function(strInfo) { $('#divMain').html(strInfo).slideToggle("slow"); LoadDiv(); }); $('#imgFeaturesEx').hide(); $('#imgFeaturesCol').show(); }); $('#imgFeaturesCol').click(function() { $('#divMain').html("").slideToggle("slow"); $('#imgFeaturesCol').hide(); $('#imgFeaturesEx').show(); }); $('#imgTechnoEx').click(function() { $.getJSON("/Visitor/GetTechnossusInfo", null, function(strInfo) { $('#divTechnossus').html(strInfo).slideToggle("slow"); }); $('#imgTechnoEx').hide(); $('#imgTechnoCol').show(); }); $('#imgTechnoCol').click(function() { $('#divTechnossus').html("").slideToggle("slow"); $('#imgTechnoCol').hide(); $('#imgTechnoEx').show(); }); })(); function LoadDiv() { $('#s4').cycle({ speed: 200, timeout: 0, pager: '#nav' }); }

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  • How do I properly add existing source code files to my Xcode project?

    - by BeachRunnerJoe
    I'm new to iPhone development and I'm still getting familiar with the Mac dev environment, including Xcode. I want to add some 3rd party code to my iPhone project, but when I add the "existing files" to my Xcode project, I'm presented with a dialog box that has far too many options that I don't understand and, as such, my project isn't working. When I #import headerfilename.h, I get a build error that reads headerfilename.h: No such file or directory. Can anyone explain to me what all these options mean or give me a link to some documentation that can? I'm having a hard time finding anything in Apple's docs. Which options do I want to choose to add existing source code files to my Xcode project? I should note that the source code files that I'm trying to add are located in my project/Classes/frameworkname/ directory. After they're added, do I need to reference this new code directory in my project settings anywhere (i.e. some kind of header file directory variable)? Thanks so much! Update: I found the following answers/responses on the apple dev forums that were very useful and helped me fix my issue... To make it simple : - if you do not check the copy option, the file stay where it is. - if you check it, it is copied in your project folders In the first case (what it seems you are doing) you need to tell the compiler that the header files are in another directory : - project info - build - search paths - User Header Search Path : add the directory from where you took the header file Hope this will help You have discovered the most confusing dialog box that ever came out of Cupertino. Six years of Xcode, and this thing still is partly a mystery to me. To even get that far, I had to make many test projects to try and reverse-engineer what this thing does. The "Copy" box means that it will copy the files as they are right now, into the project. If this box is not checked, then it just references those files during a build and copies them as they are at THAT time. For source code, you want the Copy box checked. The "relative to" is a total mystery to me and I can't help you with that. I usually leave it however it is already set. Does it mean relative to where they are on disk, or the arrangement in Xcode, or in the bundle? Who knows. The last 2 radio buttons SEEM to mean that it will either re-create the folder structure of the folder you are adding, or just put "fake" folders in Xcode that point to the real folders. This is probably your problem - you are adding source code that is not all at the top level, and when it goes to find it, it does not re-create the hierarchy. Others can supply a better way, hopefully, but what I would do is put all of the source in one folder and add that, using the Copy box. Then in Xcode you can make whatever bogus folders you want and put the source file names in those fake folders.

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  • When should I use indexed arrays of OpenGL vertices?

    - by Tartley
    I'm trying to get a clear idea of when I should be using indexed arrays of OpenGL vertices, drawn with gl[Multi]DrawElements and the like, versus when I should simply use contiguous arrays of vertices, drawn with gl[Multi]DrawArrays. (Update: The consensus in the replies I got is that one should always be using indexed vertices.) I have gone back and forth on this issue several times, so I'm going to outline my current understanding, in the hopes someone can either tell me I'm now finally more or less correct, or else point out where my remaining misunderstandings are. Specifically, I have three conclusions, in bold. Please correct them if they are wrong. One simple case is if my geometry consists of meshes to form curved surfaces. In this case, the vertices in the middle of the mesh will have identical attributes (position, normal, color, texture coord, etc) for every triangle which uses the vertex. This leads me to conclude that: 1. For geometry with few seams, indexed arrays are a big win. Follow rule 1 always, except: For geometry that is very 'blocky', in which every edge represents a seam, the benefit of indexed arrays is less obvious. To take a simple cube as an example, although each vertex is used in three different faces, we can't share vertices between them, because for a single vertex, the surface normals (and possible other things, like color and texture co-ord) will differ on each face. Hence we need to explicitly introduce redundant vertex positions into our array, so that the same position can be used several times with different normals, etc. This means that indexed arrays are of less use. e.g. When rendering a single face of a cube: 0 1 o---o |\ | | \ | | \| o---o 3 2 (this can be considered in isolation, because the seams between this face and all adjacent faces mean than none of these vertices can be shared between faces) if rendering using GL_TRIANGLE_FAN (or _STRIP), then each face of the cube can be rendered thus: verts = [v0, v1, v2, v3] colors = [c0, c0, c0, c0] normal = [n0, n0, n0, n0] Adding indices does not allow us to simplify this. From this I conclude that: 2. When rendering geometry which is all seams or mostly seams, when using GL_TRIANGLE_STRIP or _FAN, then I should never use indexed arrays, and should instead always use gl[Multi]DrawArrays. (Update: Replies indicate that this conclusion is wrong. Even though indices don't allow us to reduce the size of the arrays here, they should still be used because of other performance benefits, as discussed in the comments) The only exception to rule 2 is: When using GL_TRIANGLES (instead of strips or fans), then half of the vertices can still be re-used twice, with identical normals and colors, etc, because each cube face is rendered as two separate triangles. Again, for the same single cube face: 0 1 o---o |\ | | \ | | \| o---o 3 2 Without indices, using GL_TRIANGLES, the arrays would be something like: verts = [v0, v1, v2, v2, v3, v0] normals = [n0, n0, n0, n0, n0, n0] colors = [c0, c0, c0, c0, c0, c0] Since a vertex and a normal are often 3 floats each, and a color is often 3 bytes, that gives, for each cube face, about: verts = 6 * 3 floats = 18 floats normals = 6 * 3 floats = 18 floats colors = 6 * 3 bytes = 18 bytes = 36 floats and 18 bytes per cube face. (I understand the number of bytes might change if different types are used, the exact figures are just for illustration.) With indices, we can simplify this a little, giving: verts = [v0, v1, v2, v3] (4 * 3 = 12 floats) normals = [n0, n0, n0, n0] (4 * 3 = 12 floats) colors = [c0, c0, c0, c0] (4 * 3 = 12 bytes) indices = [0, 1, 2, 2, 3, 0] (6 shorts) = 24 floats + 12 bytes, and maybe 6 shorts, per cube face. See how in the latter case, vertices 0 and 2 are used twice, but only represented once in each of the verts, normals and colors arrays. This sounds like a small win for using indices, even in the extreme case of every single geometry edge being a seam. This leads me to conclude that: 3. When using GL_TRIANGLES, one should always use indexed arrays, even for geometry which is all seams. Please correct my conclusions in bold if they are wrong.

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  • My server app works strangely. What could be the reason(s)?

    - by Poni
    Hi! I've written a server app (two parts actually; proxy server and a game server) using C++ (board game). It uses IOCP as the sockets interface. For that app I've also written a "client simulator" (hereafter "client") app that spawns many client connections, where each of them plays, in very high speed, getting the CPU to be 100% utilized. So, that's how it goes in terms of topology: Game server - holds the game state. Real players do not connect it directly but through the proxy server. When a player joins a game, the proxy actually asks for it on behalf of that player, and the game server spawns a "player instance" for that player, and from now on, every notification between the game server and the player is being passed through the proxy. Proxy server - holds TCP connections with the real players. Players communicate with the game server through it only. Client simulator - connects to the proxy only. When running the server (again, it's actually two server apps) & client locally it all works just fine. I'm talking about 40k+ player instances in which all of them are active in a game. On the other hand, when running the server remotely with, say, 1000 clients who play things getting strange. For example, I run it as said above. Then with Task Manager I kill the client simulator app ("End Process Tree"). Then it seems like the buffer of the remote server got modified by another thread, or in other words, a memory corruption has been occurred. The server crashes because it got an unknown message id (it's a custom protocol where each message has it's own unique number). To make things clear, here is how I run the apps: PC1 - game server and clients simulator (because the clients will connect the proxy). PC2 - proxy server. The strangest thing is this: Only the remote side gets "corrupted". Remote in terms that it's not the PC I use to code the app (VC++ 2008). Let's call the PC I use to code the apps "PC1". Now for example, if this time I ran the game server on PC1 (it means that proxy server on PC2 and clients simulator on PC1), then the proxy server crashes with an "unknown message id" error. Another variation is when I run the proxy server on PC1 (again, the dev machine), the game server and the clients simulator on PC2, then the game server on PC2 gets crashed. As for the IOCP config: The servers' internal connections use the default receive/send buffer sizes. Tried even with setting them to 1MB, but no luck. I have three PCs in total; 2 x Vista 64bit <<-- one of those is the dev machine. The other is connected through WiFi. 1 x WinXP 32bit They're all connected in a "full duplex" manner. What could be the reason? Tried about everything; Stack tracing, recording some actions (like read/write logging).. I want to stress that only the PC I'm not using to code the apps crashes (actually the server app "role" which is running on it - sometimes the game server and sometimes the proxy server). At first I thought that maybe the wireless PC has problems (it's wireless..) but: TCP has it's own mechanisms to make sure the packet is delivered properly. Also, a crash also happens when trying it with the two PCs that are physically connected (Vista vs. XP). Another option is that the Windows DLLs versions might have problems, but then again, one of the tests is Vista vs. Vista, and the other is Vista vs. XP. Any idea?

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  • DSP - Filtering in the frequency domain via FFT

    - by Trap
    I've been playing around a little with the Exocortex implementation of the FFT, but I'm having some problems. Whenever I modify the amplitudes of the frequency bins before calling the iFFT the resulting signal contains some clicks and pops, especially when low frequencies are present in the signal (like drums or basses). However, this does not happen if I attenuate all the bins by the same factor. Let me put an example of the output buffer of a 4-sample FFT: // Bin 0 (DC) FFTOut[0] = 0.0000610351563 FFTOut[1] = 0.0 // Bin 1 FFTOut[2] = 0.000331878662 FFTOut[3] = 0.000629425049 // Bin 2 FFTOut[4] = -0.0000381469727 FFTOut[5] = 0.0 // Bin 3, this is the first and only negative frequency bin. FFTOut[6] = 0.000331878662 FFTOut[7] = -0.000629425049 The output is composed of pairs of floats, each representing the real and imaginay parts of a single bin. So, bin 0 (array indexes 0, 1) would represent the real and imaginary parts of the DC frequency. As you can see, bins 1 and 3 both have the same values, (except for the sign of the Im part), so I guess bin 3 is the first negative frequency, and finally indexes (4, 5) would be the last positive frequency bin. Then to attenuate the frequency bin 1 this is what I do: // Attenuate the 'positive' bin FFTOut[2] *= 0.5; FFTOut[3] *= 0.5; // Attenuate its corresponding negative bin. FFTOut[6] *= 0.5; FFTOut[7] *= 0.5; For the actual tests I'm using a 1024-length FFT and I always provide all the samples so no 0-padding is needed. // Attenuate var halfSize = fftWindowLength / 2; float leftFreq = 0f; float rightFreq = 22050f; for( var c = 1; c < halfSize; c++ ) { var freq = c * (44100d / halfSize); // Calc. positive and negative frequency indexes. var k = c * 2; var nk = (fftWindowLength - c) * 2; // This kind of attenuation corresponds to a high-pass filter. // The attenuation at the transition band is linearly applied, could // this be the cause of the distortion of low frequencies? var attn = (freq < leftFreq) ? 0 : (freq < rightFreq) ? ((freq - leftFreq) / (rightFreq - leftFreq)) : 1; // Attenuate positive and negative bins. mFFTOut[ k ] *= (float)attn; mFFTOut[ k + 1 ] *= (float)attn; mFFTOut[ nk ] *= (float)attn; mFFTOut[ nk + 1 ] *= (float)attn; } Obviously I'm doing something wrong but can't figure out what. I don't want to use the FFT output as a means to generate a set of FIR coefficients since I'm trying to implement a very basic dynamic equalizer. What's the correct way to filter in the frequency domain? what I'm missing? Thanks in advance.

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  • Boost::Spirit::Qi autorules -- avoiding repeated copying of AST data structures

    - by phooji
    I've been using Qi and Karma to do some processing on several small languages. Most of the grammars are pretty small (20-40 rules). I've been able to use autorules almost exclusively, so my parse trees consist entirely of variants, structs, and std::vectors. This setup works great for the common case: 1) parse something (Qi), 2) make minor manipulations to the parse tree (visitor), and 3) output something (Karma). However, I'm concerned about what will happen if I want to make complex structural changes to a syntax tree, like moving big subtrees around. Consider the following toy example: A grammar for s-expr-style logical expressions that uses autorules... // Inside grammar class; rule names match struct names... pexpr %= pand | por | var | bconst; pand %= lit("(and ") >> (pexpr % lit(" ")) >> ")"; por %= lit("(or ") >> (pexpr % lit(" ")) >> ")"; pnot %= lit("(not ") >> pexpr >> ")"; ... which leads to parse tree representation that looks like this... struct var { std::string name; }; struct bconst { bool val; }; struct pand; struct por; struct pnot; typedef boost::variant<bconst, var, boost::recursive_wrapper<pand>, boost::recursive_wrapper<por>, boost::recursive_wrapper<pnot> > pexpr; struct pand { std::vector<pexpr> operands; }; struct por { std::vector<pexpr> operands; }; struct pnot { pexpr victim; }; // Many Fusion Macros here Suppose I have a parse tree that looks something like this: pand / ... \ por por / \ / \ var var var var (The ellipsis means 'many more children of similar shape for pand.') Now, suppose that I want negate each of the por nodes, so that the end result is: pand / ... \ pnot pnot | | por por / \ / \ var var var var The direct approach would be, for each por subtree: - create pnot node (copies por in construction); - re-assign the appropriate vector slot in the pand node (copies pnot node and its por subtree). Alternatively, I could construct a separate vector, and then replace (swap) the pand vector wholesale, eliminating a second round of copying. All of this seems cumbersome compared to a pointer-based tree representation, which would allow for the pnot nodes to be inserted without any copying of existing nodes. My question: Is there a way to avoid copy-heavy tree manipulations with autorule-compliant data structures? Should I bite the bullet and just use non-autorules to build a pointer-based AST (e.g., http://boost-spirit.com/home/2010/03/11/s-expressions-and-variants/)?

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  • Help with specific Regex: need to match multiple instances of multiple formats in a single string.

    - by KevenK
    I apologize for the terrible title...it can be hard to try to summarize an entire situation into a single sentence. Let me start by saying that I'm asking because I'm just not a Regex expert. I've used it a bit here and there, but I just come up short with the correct way to meet the following requirements. The Regex that I'm attempting to write is for use in an XML schema for input validation, and used elsewhere in Javascript for the same purpose. There are two different possible formats that are supported. There is a literal string, which must be surrounded by quotation marks, and a Hex-value string which must be surrounded by braces. Some test cases: "this is a literal string" <-- Valid string, enclosed properly in "s "this should " still be correct" <-- Valid string, "s are allowed within (if possible, this requirement could be forgiven if necessary) "{00 11 22}" <-- Valid string, {}'s allow in strings. Another one that can be forgiven if necessary I am bad output <-- Invalid string, no "s "Some more problemss"you know <-- Invalid string, must be fully contained in "s {0A 68 4F 89 AC D2} <-- Valid string, hex characters enclosed in {}s {DDFF1234} <-- Valid string, spaces are ignored for Hex strings DEADBEEF <-- Invalid string, must be contained in either "s or {}s {0A 12 ZZ} <-- Invalid string, 'Z' is not a valid Hex character To satisfy these general requirements, I had come up with the following Regex that seems to work well enough. I'm still fairly new to Regex, so there could be a huge hole here that I'm missing.: &quot;.+&quot;|\{([0-9]|[a-f]|[A-F]| )+\} If I recall correctly, the XML Schema regex automatically assumes beginning and end of line (^ and $ respectively). So, essentially, this regex accepts any string that starts and ends with a ", or starts and ends with {}s and contains only valid Hexidecimal characters. This has worked well for me so far except that I had forgotten about another (although less common, and thus forgotten) input option that completely breaks my regex. Where I made my mistake: Valid input should also allow a user to separate valid strings (of either type, literal/hex) by a comma. This means that a single string should be able to contain more than one of the above valid strings, separated by commas. Luckily, however, a comma is not a supported character within a literal string (although I see that my existing regex does not care about commas). Example test cases: "some string",{0A F1} <-- Valid {1122},{face},"peanut butter" <-- Valid {0D 0A FF FE},"string",{FF FFAC19 85} <-- Valid (Spaces don't matter in Hex values) "Validation is allowed to break, if a comma is found not separating values",{0d 0a} <-- Invalid, comma is a delimiter, but "Validation is allowed to break" and "if a comma..." are not marked as separate strings with "s hi mom,"hello" <-- Invalid, String1 was not enclosed properly in "s or {}s My thoughts are that it is possible to use commas as a delimiter to check each "section" of the string to match a regex similar to the original, but I just am not that advanced in regex yet to come up with a solution on my own. Any help would be appreciated, but ultimately a final solution with an explanation would just stellar. Thanks for reading this huge wall of text!

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  • How to Force an Exception from a Task to be Observed in a Continuation Task?

    - by Richard
    I have a task to perform an HttpWebRequest using Task<WebResponse>.Factory.FromAsync(req.BeginGetRespone, req.EndGetResponse) which can obviously fail with a WebException. To the caller I want to return a Task<HttpResult> where HttpResult is a helper type to encapsulate the response (or not). In this case a 4xx or 5xx response is not an exception. Therefore I've attached two continuations to the request task. One with TaskContinuationOptions OnlyOnRanToCompletion and the other with OnlyOnOnFaulted. And then wrapped the whole thing in a Task<HttpResult> to pick up the one result whichever continuation completes. Each of the three child tasks (request plus two continuations) is created with the AttachedToParent option. But when the caller waits on the returned outer task, an AggregateException is thrown is the request failed. I want to, in the on faulted continuation, observe the WebException so the client code can just look at the result. Adding a Wait in the on fault continuation throws, but a try-catch around this doesn't help. Nor does looking at the Exception property (as section "Observing Exceptions By Using the Task.Exception Property" hints here). I could install a UnobservedTaskException event handler to filter, but as the event offers no direct link to the faulted task this will likely interact outside this part of the application and is a case of a sledgehammer to crack a nut. Given an instance of a faulted Task<T> is there any means of flagging it as "fault handled"? Simplified code: public static Task<HttpResult> Start(Uri url) { var webReq = BuildHttpWebRequest(url); var result = new HttpResult(); var taskOuter = Task<HttpResult>.Factory.StartNew(() => { var tRequest = Task<WebResponse>.Factory.FromAsync( webReq.BeginGetResponse, webReq.EndGetResponse, null, TaskCreationOptions.AttachedToParent); var tError = tRequest.ContinueWith<HttpResult>( t => HandleWebRequestError(t, result), TaskContinuationOptions.AttachedToParent |TaskContinuationOptions.OnlyOnFaulted); var tSuccess = tRequest.ContinueWith<HttpResult>( t => HandleWebRequestSuccess(t, result), TaskContinuationOptions.AttachedToParent |TaskContinuationOptions.OnlyOnRanToCompletion); return result; }); return taskOuter; } with: private static HttpDownloaderResult HandleWebRequestError( Task<WebResponse> respTask, HttpResult result) { Debug.Assert(respTask.Status == TaskStatus.Faulted); Debug.Assert(respTask.Exception.InnerException is WebException); // Try and observe the fault: Doesn't help. try { respTask.Wait(); } catch (AggregateException e) { Log("HandleWebRequestError: waiting on antecedent task threw inner: " + e.InnerException.Message); } // ... populate result with details of the failure for the client ... return result; } (HandleWebRequestSuccess will eventually spin off further tasks to get the content of the response...) The client should be able to wait on the task and then look at its result, without it throwing due to a fault that is expected and already handled.

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  • Adding to database with multiple text boxes

    - by kira423
    What I am trying to do with this script is allow users to update a url for their websites, and since each user isn't going to have the same amount of websites is is hard for me to just add $_POST['website'] for each of these. Here is the script <?php include("config.php"); include("header.php"); include("functions.php"); if(!isset($_SESSION['username']) && !isset($_SESSION['password'])){ header("Location: pubs.php"); } $getmember = mysql_query("SELECT * FROM `publishers` WHERE username = '".$_SESSION['username']."'"); $info = mysql_fetch_array($getmember); $getsites = mysql_query("SELECT * FROM `websites` WHERE publisher = '".$info['username']."'"); $postback = $_POST['website']; $webname = $_POST['webid']; if($_POST['submit']){ var_dump( $_POST['website'] ); $update = mysql_query("UPDATE `websites` SET `postback` = '$postback' WHERE name = '$webname'"); } print" <div id='center'> <span id='tools_lander'><a href='export.php'>Export Campaigns</a></span> <div id='calendar_holder'> <h3>Please define a postback for each of your websites below. The following variables should be used when creating your postback.<br /> cid = Campaign ID<br /> sid = Sub ID<br /> rate = Campaign Rate<br /> status = Status of Lead. 1 means payable 2 mean reversed<br /> A sample postback URL would be <br /> http://www.example.com/postback.php?cid=#cid&sid=#sid&rate=#rate&status=#status</h3> <table class='balances' align='center'> <form method='POST' action=''>"; while($website = mysql_fetch_array($getsites)){ print" <tr> <input type ='hidden' name='webid' value='".$website['id']."' /> <td style='font-weight:bold;'>".$website['name']."'s Postback:</td> <td><input type='text' style='width:400px;' name='website[]' value='".$website['postback']."' /></td> </tr>"; } print" <td style='float:right;position:relative;left:150px;'><input type='submit' name='submit' style='font-size:15px;height:30px;width:100px;' value='Submit' /></td> </form> </table> </div>"; include("footer.php"); ?> What I am attempting to do insert the what is inputted in the text boxes to their corresponding websites, and I cannot think of any other way to do it, and this obviously does not works and returns a notice stating Array to string conversion If there is a more logical way to do this please let me know.

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  • Broken corba object references

    - by cube
    I'm working on a homework and got stuck. The task is to serve objects using a default servant. But when I try to use the reference, weird things happen. Some part of corba prints a stack trace, but no exception is thrown. The problem happens when the server receives the reference and should call some method on it. The reference is then shortened and doesn't contain the object ID (which means that my servant implementation can't do anything reasonable). This is the implementation of the servant, where the problem appears: public class ModelFileImpl extends ModelFilePOA{ @Override public String getName() { try { return new String(_poa().reference_to_id(_this_object())); } catch (Throwable e) {} assert false; return null; } } If I take _this_object().toString() inside the try block and put it into dior -i i get this: ------IOR components----- TypeId : IDL:termproject/idl/ModelFile:1.0 TAG_INTERNET_IOP Profiles: Profile Id: 0 IIOP Version: 1.2 Host: 127.0.0.1 Port: 45954 Object key (URL): %AF%AB%CB%00%00%00%00%20Q%BA%F4%FF%00%00%00%01%00%00%00%00%00%00%00%01%0000%00%08RootPOA%00%00%00%00%08%00%00%00%02%00%00%00%00%14 Object key (hex): 0xAF AB CB 00 00 00 00 20 51 BA F4 FF 00 00 00 01 00 00 00 00 00 00 00 01 00 00 00 08 52 6F 6F 74 50 4F 41 00 00 00 00 08 00 00 00 02 00 00 00 00 14 -- Found 2 Tagged Components-- #0: TAG_CODE_SETS ForChar native code set Id: ISO8859_1 Char Conversion Code Sets: UTF8 , Unknown TCS: 10020 ForWChar native code set Id: UTF16 WChar Conversion Code Sets: Unknown TCS: 10100 Unknown tag : 38 however the part of server that makes the reference and the client see the reference as ------IOR components----- TypeId : IDL:termproject/idl/ModelFile:1.0 TAG_INTERNET_IOP Profiles: Profile Id: 0 IIOP Version: 1.2 Host: 127.0.0.1 Port: 45954 Object key (URL): %AF%AB%CB%00%00%00%00%20Q%BA%F4%FF%00%00%00%01%00%00%00%00%00%00%00%02%00%00%00%08RootPOA%00%00%00%00%09modelPoa%00%00%00%00%00%00%00%10testModel1.MyIDL%14 Object key (hex): 0xAF AB CB 00 00 00 00 20 51 BA F4 FF 00 00 00 01 00 00 00 00 00 00 00 02 00 00 00 08 52 6F 6F 74 50 4F 41 00 00 00 00 09 6D 6F 64 65 6C 50 6F 61 00 00 00 00 00 00 00 10 74 65 73 74 4D 6F 64 65 6C 31 2E 4D 79 49 44 4C 14 -- Found 2 Tagged Components-- #0: TAG_CODE_SETS ForChar native code set Id: ISO8859_1 Char Conversion Code Sets: UTF8 , Unknown TCS: 10020 ForWChar native code set Id: UTF16 WChar Conversion Code Sets: Unknown TCS: 10100 Unknown tag : 38 ("modelPoa" (the name of the poa working with default clients) and "testModel1.MyIDL" (the identifier of the object) in the object key are missing in the first one) I've tried sniffing the traffic and found out that the client still sends the correct reference. This is how i create the references: ret[i] = ModelFileHelper.narrow(modelFilePoa.create_reference_with_id(files[i].getBytes(), ModelFileHelper.id())); And this is how i set up the server: // init ORB ORB orb = ORB.init(args, null); // init POA POA poa = POAHelper.narrow(orb.resolve_initial_references("RootPOA")); // create the POA for the models. Policy[] policies = { poa.create_request_processing_policy(RequestProcessingPolicyValue.USE_DEFAULT_SERVANT), poa.create_servant_retention_policy(ServantRetentionPolicyValue.NON_RETAIN), poa.create_id_assignment_policy(IdAssignmentPolicyValue.USER_ID) }; POA modelPoa = poa.create_POA("modelPoa", poa.the_POAManager(), policies); modelPoa.the_POAManager().activate(); modelPoa.set_servant(new ModelFileImpl()); modelPoa.the_POAManager().activate(); ModelStoreImpl impl = new ModelStoreImpl(modelPoa); // create the object reference org.omg.CORBA.Object obj = poa.servant_to_reference(impl); // ... store the IOR file ... orb.run(); I'd be really grateful for any pointers (or references :-) )

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  • Doxygen for C++ template class member specialization

    - by Ziv
    When I write class templates, and need to fully-specialize members of those classes, Doxygen doesn't recognize the specialization - it documents only the generic definition, or (if there are only specializations) the last definition. Here's a simple example: ===MyClass.hpp=== #ifndef MYCLASS_HPP #define MYCLASS_HPP template<class T> class MyClass{ public: static void foo(); static const int INT_CONST; static const T TTYPE_CONST; }; /* generic definitions */ template<class T> void MyClass<T>::foo(){ printf("Generic foo\n"); } template<class T> const int MyClass<T>::INT_CONST = 5; /* specialization declarations */ template<> void MyClass<double>::foo(); template<> const int MyClass<double>::INT_CONST; template<> const double MyClass<double>::TTYPE_CONST; template<> const char MyClass<char>::TTYPE_CONST; #endif === MyClass.cpp === #include "MyClass.hpp" /* specialization definitions */ template<> void MyClass<double>::foo(){ printf("Specialized double foo\n"); } template<> const int MyClass<double>::INT_CONST = 10; template<> const double MyClass<double>::TTYPE_CONST = 3.141; template<> const char MyClass<char>::TTYPE_CONST = 'a'; So in this case, foo() will be documented as printing "Generic foo," INT_CONST will be documented as set to 5, with no mention of the specializations, and TTYPE_CONST will be documented as set to 'a', with no mention of 3.141 and no indication that 'a' is a specialized case. I need to be able to document the specializations - either within the documentation for MyClass<T>, or on new pages for MyClass<double>, MyClass<char>. How do I do this? Can Doxygen even handle this? Am I possibly doing something wrong in the declarations/code structure that's keeping Doxygen from understanding what I want? I should note two related cases: A) For templated functions, specialization works fine, e.g.: /* functions that are global/in a namespace */ template<class T> void foo(){ printf("Generic foo\n"); } template<> void foo<double>(){ printf("Specialized double foo\n"); } This will document both foo() and foo(). B) If I redeclare the entire template, i.e. template<> class MyClass<double>{...};, then MyClass<double> will get its own documentation page, as a seperate class. But this means actually declaring an entirely new class - there is no relation between MyClass<T> and MyClass<double> if MyClass<double> itself is declared. So I'd have to redeclare the class and all its members, and repeat all the definitions of class members, specialized for MyClass<double>, all to make it appear as though they're using the same template. Very awkward, feels like a kludge solution. Suggestions? Thanks much :) --Ziv

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  • tkinter frame does not show on startup

    - by Jzz
    this is my first question on SO, so correct me please if I make a fool of myself. I have this fairly complicated python / Tkinter application (python 2.7). On startup, the __init__ loads several frames, and loads a database. When that is finished, I want to set the application to a default state (there are 2 program states, 'calculate' and 'config'). Setting the state of the application means that the appropriate frame is displayed (using grid). When the program is running, the user can select a program state in the menu. Problem is, the frame is not displayed on startup. I get an empty application (menu bar and status bar are displayed). When I select a program state in the menu, the frame displays as it should. Question: What am I doing wrong? Should I update idletasks? I tried, but no result. Anything else? Background: I use the following to switch program states: def set_program_state(self, state): '''sets the program state''' #try cleaning all the frames: try: self.config_frame.grid_forget() except: pass try: self.tidal_calculations_frame.grid_forget() except: pass try: self.tidal_grapth_frame.grid_forget() except: pass if state == "calculate": print "Switching to calculation mode" self.tidal_calculations_frame.grid() #frame is preloaded self.tidal_calculations_frame.fill_data(routes=self.routing_data.routes, deviations=self.misc_data.deviations, ship_types=self.misc_data.ship_types) self.tidal_grapth_frame.grid() self.program_state = "calculate" elif state == "config": print "Switching to config mode" self.config_frame = GUI_helper.config_screen_frame(self, self.user) #load frame first (contents depend on type of user) self.config_frame.grid() self.program_state = "config" I understand that this is kind of messy to read, so I simplified things for testing, using this: def set_program_state(self, state): '''sets the program state''' #try cleaning all the frames: try: self.testlabel_1.grid_forget() except: pass try: self.testlabel_2.grid_forget() except: pass if state == "calculate": print "switching to test1" self.testlabel_1 = tk.Label(self, text="calculate", borderwidth=1, relief=tk.RAISED) self.testlabel_1.grid(row=0, sticky=tk.W+tk.E) elif state == "config": print "switching to test1" self.testlabel_2 = tk.Label(self, text="config", borderwidth=1, relief=tk.RAISED) self.testlabel_2.grid(row=0, sticky=tk.W+tk.E) But the result is the same. The frame (or label in this test) is not displayed at startup, but when the user selects the state (calling the same function) the frame is displayed. UPDATE the sample code in the comments (thanks for that!) pointed me in another direction. Further testing revealed (what I think) the cause of the problem. Disabling the display of the status bar made the program work as expected. Turns out, I used pack to display the statusbar and grid to display the frames. And they are in the same container, so problems arise. I fixed that by using only pack inside the main container. But the same problem is still there. This is what I use for the statusbar: self.status = GUI_helper.StatusBar(self.parent) self.status.pack(side=tk.BOTTOM, fill=tk.X) And if I comment out the last line (pack), the config frame loads on startup, as per this line: self.set_program_state("config") But if I let the status bar pack inside the main window, the config frame does not show. Where it does show when the user asks for it (with the same command as above).

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  • Why does this Quicksort work?

    - by IVlad
    I find this Quicksort partitioning approach confusing and wrong, yet it seems to work. I am referring to this pseudocode. Note: they also have a C implementation at the end of the article, but it's very different from their pseudocode, so I don't care about that. I have also written it in C like this, trying to stay true to the pseudocode as much as possible, even if that means doing some weird C stuff: #include <stdio.h> int partition(int a[], int p, int r) { int x = a[p]; int i = p - 1; int j = r + 1; while (1) { do j = j - 1; while (!(a[j] <= x)); do i = i + 1; while (!(a[i] >= x)); if (i < j) { int t = a[i]; a[i] = a[j]; a[j] = t; } else { for (i = 1; i <= a[0]; ++i) printf("%d ", a[i]); printf("- %d\n", j); return j; } } } int main() { int a[100] = //{8, 6,10,13,15,8,3,2,12}; {7, 7, 6, 2, 3, 8, 4, 1}; partition(a, 1, a[0]); return 0; } If you run this, you'll get the following output: 1 6 2 3 4 8 7 - 5 However, this is wrong, isn't it? Clearly a[5] does not have all the values before it lower than it, since a[2] = 6 > a[5] = 4. Not to mention that 7 is supposed to be the pivot (the initial a[p]) and yet its position is both incorrect and lost. The following partition algorithm is taken from wikipedia: int partition2(int a[], int p, int r) { int x = a[r]; int store = p; for (int i = p; i < r; ++i) { if (a[i] <= x) { int t = a[i]; a[i] = a[store]; a[store] = t; ++store; } } int t = a[r]; a[r] = a[store]; a[store] = t; for (int i = 1; i <= a[0]; ++i) printf("%d ", a[i]); printf("- %d\n", store); return store; } And produces this output: 1 6 2 3 8 4 7 - 1 Which is a correct result in my opinion: the pivot (a[r] = a[7]) has reached its final position. However, if I use the initial partitioning function in the following algorithm: void Quicksort(int a[], int p, int r) { if (p < r) { int q = partition(a, p, r); // initial partitioning function Quicksort(a, p, q); Quicksort(a, q + 1, r); // I'm pretty sure q + r was a typo, it doesn't work with q + r. } } ... it seems to be a correct sorting algorithm. I tested it out on a lot of random inputs, including all 0-1 arrays of length 20. I have also tried using this partition function for a selection algorithm, in which it failed to produce correct results. It seems to work and it's even very fast as part of the quicksort algorithm however. So my questions are: Can anyone post an example on which the algorithm DOESN'T work? If not, why does it work, since the partitioning part seems to be wrong? Is this another partitioning approach that I don't know about?

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  • Picking good first estimates for Goldschmidt division

    - by Mads Elvheim
    I'm calculating fixedpoint reciprocals in Q22.10 with Goldschmidt division for use in my software rasterizer on ARM. This is done by just setting the nominator to 1, i.e the nominator becomes the scalar on the first iteration. To be honest, I'm kind of following the wikipedia algorithm blindly here. The article says that if the denominator is scaled in the half-open range (0.5, 1.0], a good first estimate can be based on the denominator alone: Let F be the estimated scalar and D be the denominator, then F = 2 - D. But when doing this, I lose a lot of precision. Say if I want to find the reciprocal of 512.00002f. In order to scale the number down, I lose 10 bits of precision in the fraction part, which is shifted out. So, my questions are: Is there a way to pick a better estimate which does not require normalization? Also, is it possible to pre-calculate the first estimates so the series converges faster? Right now, it converges after the 4th iteration on average. On ARM this is about ~50 cycles worst case, and that's not taking emulation of clz/bsr into account, nor memory lookups. Here is my testcase. Note: The software implementation of clz on line 13 is from my post here. You can replace it with an intrinsic if you want. #include <stdio.h> #include <stdint.h> const unsigned int BASE = 22ULL; static unsigned int divfp(unsigned int val, int* iter) { /* Nominator, denominator, estimate scalar and previous denominator */ unsigned long long N,D,F, DPREV; int bitpos; *iter = 1; D = val; /* Get the shift amount + is right-shift, - is left-shift. */ bitpos = 31 - clz(val) - BASE; /* Normalize into the half-range (0.5, 1.0] */ if(0 < bitpos) D >>= bitpos; else D <<= (-bitpos); /* (FNi / FDi) == (FN(i+1) / FD(i+1)) */ /* F = 2 - D */ F = (2ULL<<BASE) - D; /* N = F for the first iteration, because the nominator is simply 1. So don't waste a 64-bit UMULL on a multiply with 1 */ N = F; D = ((unsigned long long)D*F)>>BASE; while(1){ DPREV = D; F = (2<<(BASE)) - D; D = ((unsigned long long)D*F)>>BASE; /* Bail when we get the same value for two denominators in a row. This means that the error is too small to make any further progress. */ if(D == DPREV) break; N = ((unsigned long long)N*F)>>BASE; *iter = *iter + 1; } if(0 < bitpos) N >>= bitpos; else N <<= (-bitpos); return N; } int main(int argc, char* argv[]) { double fv, fa; int iter; unsigned int D, result; sscanf(argv[1], "%lf", &fv); D = fv*(double)(1<<BASE); result = divfp(D, &iter); fa = (double)result / (double)(1UL << BASE); printf("Value: %8.8lf 1/value: %8.8lf FP value: 0x%.8X\n", fv, fa, result); printf("iteration: %d\n",iter); return 0; }

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  • Decentralized synchronized secure data storage

    - by Alberich
    Introduction Hi, I am going to ask a question which seems utopic for me, but I need to know if there is a way to achieve what I need. And if not, I need to know why not. The idea Suppose I have a database structure, in MySql. I want to create some solution to allow anyone (no matter who, no matter where) to have a synchronized copy (updated clone) of this database (with its content) Well, and it is not going to be just one synchronized copy, it could (and should) be a multiple replication (supposing the basic, this means, for example, ten copies all over the world) And, the most important thing: It must be secure. By secure I mean only real-accepted transactions will be synchronized with all the others (no matter how many) database copies/clones. Note: Since it would be quite difficult to make the synchronization in real-time, I will design everything to make this feature dispensable. So it is not required. My auto-suggestion This is how I am thinking to manage it: Time identifiers and Updates checking: Every action (insert, update, delete...) will be stored as the action instruction itself, associated to the time identifier. [I think better than a DATETIME field, it'll be an INT one, with the number of miliseconds passed from 1st january 2013 on, for example]. So each copy is going to ask to the "neighbour copy" for new actions done since last update, and execute them after checking they are allowed. Problem 1: the "neighbour copy" could be outdated too. Solution 1: do not ask just one neighbour, create a random list with some of the copies/clones and ask them for news (I could avoid the list and ask ALL the clones for updates, but this will be inefficient if clones number ascends too much). Problem 2: Real-time global synchronization is not active. What if... Someone at CLONE_ENTERPRISING inserts a row into TABLE. ... this row goes to every clone ... Someone at CLONE_FIXEMALL deletes this row. ... and at the same time, somewhere in an outdated clone ... Someone at CLONE_DROPOUT edits this row (now inexistent at the other clones) Solution 2: easy stuff, force a GLOBAL synchronization before doing any new "depending-on-third-data action" (edit, for example). This global synch. will be unnecessary when making an INSERT, for instance. Note: Well, someone could have some fun, and make the same insert in two clones... since they're not getting updated in real-time, this row will exist twice. But, it's the same as when we have one single database, in some needed cases we check if there is an existing same-row before doing the final action. Not a problem. Problem 3: It is possible to edit the code and do not filter actions, so someone could spread instructions to delete everything, or just make some trolling activity. This is not a problem, since good clones will always be somewhere. Those who got bad won't interest anymore. I really appreciate if you read. I know this is not the perfect solution, it has possibly hundred of holes, but it is my basic start. I will now appreciate anything you can teach me now. Thanks a lot. PS.: It could be that all this I am trying already exists and has its own name. Sorry for asking then (I'd anyway thank this name, if it exists)

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  • using getScript to import plugin on page using multiple versions of jQuery

    - by mikez302
    I am developing an app on a page that uses jQuery 1.2.6, but I would like to use jQuery 1.4.2 for my app. I really don't like to use multiple versions of jQuery like this but the copy on the page (1.2.6) is something I have no control over. I decided to isolate my code like this: <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html><head> <script type="text/javascript" src="jquery-1.2.6.min.js> <script type="text/javascript" src="pageStuff.js"> </head> <body> Welcome to our page. <div id="app"> <script type="text/javascript" src="http://ajax.googleapis.com/ajax/libs/jquery/1.4.2/jquery.js"></script> <script type="text/javascript" src="myStuff.js"> </div> </body></html> The file myStuff.js has my own code that is supposed to use jQuery 1.4.2, and it looks like this: (function($) { //wrap everything in function to add ability to use $ var with noConflict var jQuery = $; //my code })(jQuery.noConflict(true)); This is an extremely simplified version, but I hope you get the idea of what I did. For a while, everything worked fine. However, I decided to want to use a jQuery plugin in a separate file. I tested it and it acted funny. After some experimentation, I found out that the plugin was using the old version of jQuery, when I wanted it to use the new version. Does anyone know how to import and run a js file from the context within the function wrapping the code in myStuff.js? In case this matters to anyone, here is how I know the plugin is using the old version, and what I did to try to solve the problem: I made a file called test.js, consisting of this line: alert($.fn.jquery); I tried referencing the file in a script tag the way external Javascript is usually included, below myStuff.js, and it came up as 1.2.6, like I expected. I then got rid of that script tag and put this line in myStuff.js: $.getScript("test.js"); and it still came back as 1.2.6. That wasn't a big surprise -- according to jQuery's documentation, scripts included that way are executed in the global context. I then tried doing this instead: var testFn = $.proxy($.getScript, this); testFn("test.js"); and it still came back as 1.2.6. After some tinkering, I found out that the "this" keyword referred to the window, which I assume means the global context. I am looking for something to put in place of "this" to refer to the context of the enclosing function, or some other way to make the code in the file run from the enclosing function. I noticed that if I copy and paste the code, it works fine, but it is a big plugin that is used in many places, and I would prefer not to clutter up my file with their code. I am out of ideas. Does anyone else know how to do this?

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  • Do you have suggestions for these assembly mnemonics?

    - by Noctis Skytower
    Greetings! Last semester in college, my teacher in the Computer Languages class taught us the esoteric language named Whitespace. In the interest of learning the language better with a very busy schedule (midterms), I wrote an interpreter and assembler in Python. An assembly language was designed to facilitate writing programs easily, and a sample program was written with the given assembly mnemonics. Now that it is summer, a new project has begun with the objective being to rewrite the interpreter and assembler for Whitespace 0.3, with further developments coming afterwards. Since there is so much extra time than before to work on its design, you are presented here with an outline that provides a revised set of mnemonics for the assembly language. This post is marked as a wiki for their discussion. Have you ever had any experience with assembly languages in the past? Were there some instructions that you thought should have been renamed to something different? Did you find yourself thinking outside the box and with a different paradigm than in which the mnemonics were named? If you can answer yes to any of those questions, you are most welcome here. Subjective answers are appreciated! Stack Manipulation (IMP: [Space]) Stack manipulation is one of the more common operations, hence the shortness of the IMP [Space]. There are four stack instructions. hold N Push the number onto the stack copy Duplicate the top item on the stack copy N Copy the nth item on the stack (given by the argument) onto the top of the stack swap Swap the top two items on the stack drop Discard the top item on the stack drop N Slide n items off the stack, keeping the top item Arithmetic (IMP: [Tab][Space]) Arithmetic commands operate on the top two items on the stack, and replace them with the result of the operation. The first item pushed is considered to be left of the operator. add Addition sub Subtraction mul Multiplication div Integer Division mod Modulo Heap Access (IMP: [Tab][Tab]) Heap access commands look at the stack to find the address of items to be stored or retrieved. To store an item, push the address then the value and run the store command. To retrieve an item, push the address and run the retrieve command, which will place the value stored in the location at the top of the stack. save Store load Retrieve Flow Control (IMP: [LF]) Flow control operations are also common. Subroutines are marked by labels, as well as the targets of conditional and unconditional jumps, by which loops can be implemented. Programs must be ended by means of [LF][LF][LF] so that the interpreter can exit cleanly. L: Mark a location in the program call L Call a subroutine goto L Jump unconditionally to a label if=0 L Jump to a label if the top of the stack is zero if<0 L Jump to a label if the top of the stack is negative return End a subroutine and transfer control back to the caller halt End the program I/O (IMP: [Tab][LF]) Finally, we need to be able to interact with the user. There are IO instructions for reading and writing numbers and individual characters. With these, string manipulation routines can be written. The read instructions take the heap address in which to store the result from the top of the stack. print chr Output the character at the top of the stack print int Output the number at the top of the stack input chr Read a character and place it in the location given by the top of the stack input int Read a number and place it in the location given by the top of the stack Question: How would you redesign, rewrite, or rename the previous mnemonics and for what reasons?

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  • What is there so useful in the Decorator Pattern? My example doesn't work

    - by Green
    The book says: The decorator pattern can be used to extend (decorate) the functionality of a certain object I have a rabbit animal. And I want my rabbit to have, for example, reptile skin. Just want to decorate a common rabbit with reptile skin. I have the code. First I have abstract class Animal with everythig that is common to any animal: abstract class Animal { abstract public function setSleep($hours); abstract public function setEat($food); abstract public function getSkinType(); /* and more methods which for sure will be implemented in any concrete animal */ } I create class for my rabbit: class Rabbit extends Animal { private $rest; private $stomach; private $skinType = "hair"; public function setSleep($hours) { $this->rest = $hours; } public function setFood($food) { $this->stomach = $food; } public function getSkinType() { return $this->$skinType; } } Up to now everything is OK. Then I create abstract AnimalDecorator class which extends Animal: abstract class AnimalDecorator extends Animal { protected $animal; public function __construct(Animal $animal) { $this->animal = $animal; } } And here the problem comes. Pay attention that AnimalDecorator also gets all the abstract methods from the Animal class (in this example just two but in real can have many more). Then I create concrete ReptileSkinDecorator class which extends AnimalDecorator. It also has those the same two abstract methods from Animal: class ReptileSkinDecorator extends AnimalDecorator { public function getSkinColor() { $skin = $this->animal->getSkinType(); $skin = "reptile"; return $skin; } } And finaly I want to decorate my rabbit with reptile skin: $reptileSkinRabbit = ReptileSkinDecorator(new Rabbit()); But I can't do this because I have two abstract methods in ReptileSkinDecorator class. They are: abstract public function setSleep($hours); abstract public function setEat($food); So, instead of just re-decorating only skin I also have to re-decorate setSleep() and setEat(); methods. But I don't need to. In all the book examples there is always ONLY ONE abstract method in Animal class. And of course it works then. But here I just made very simple real life example and tried to use the Decorator pattern and it doesn't work without implementing those abstract methods in ReptileSkinDecorator class. It means that if I want to use my example I have to create a brand new rabbit and implement for it its own setSleep() and setEat() methods. OK, let it be. But then this brand new rabbit has the instance of commont Rabbit I passed to ReptileSkinDecorator: $reptileSkinRabbit = ReptileSkinDecorator(new Rabbit()); I have one common rabbit instance with its own methods in the reptileSkinRabbit instance which in its turn has its own reptileSkinRabbit methods. I have rabbit in rabbit. But I think I don't have to have such possibility. I don't understand the Decarator pattern right way. Kindly ask you to point on any mistakes in my example, in my understanding of this pattern. Thank you.

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  • Google Chrome Frame and Facebook Javascript SDK - Cannot login

    - by Giannis Savvakis
    On the example below i have an html page with the javascript code needed to login to facebook. On the i have the Google Chrome Frame meta tag that makes the page run with google chrome frame. If you open this page with any browser the finish() callback runs normally. If you open it with Google Chrome Frame it never fires. So this means that every Facebook App that tries to login to gather user data cannot login. This happens if the page is opened with google frame. But even if i remove the meta tag so that the page can open with IE8 the page opens again with google chrome frame because Facebook opens google chrome frame by default. So because this is a Facebook app that runs inside an inside facebook.com it is forced to open with Google Chrome Frame! SERIOUS BUG! I have seen other people reporting it, someone has made a test facebook app also here: http://apps.facebook.com/gcftest/ appID and channelUrl are dummy in the example below. <html xmlns="http://www.w3.org/1999/xhtml" xmlns:fb="http://www.facebook.com/2008/fbml"> <head> <meta name="generator" content= "HTML Tidy for Linux/x86 (vers 11 February 2007), see www.w3.org" /> <meta charset="utf-8" /> <meta http-equiv="Cache-Control" content="no-cache, no-store, must-revalidate" /> <meta http-equiv="Pragma" content="no-cache" /> <meta http-equiv="Expires" content="0" /> <meta http-equiv="X-UA-Compatible" content="IE=Edge,chrome=IE8" /> <title>Facebook Login</title> <script type="text/javascript"> //<![CDATA[ // Load the SDK Asynchronously (function(d){ var js, id = 'facebook-jssdk', ref = d.getElementsByTagName('script')[0]; if (d.getElementById(id)) { return; } js = d.createElement('script'); js.id = id; js.async = true; js.src = "//connect.facebook.net/en_US/all.js"; ref.parentNode.insertBefore(js, ref); }(document)); var appID = '0000000000000'; var channelUrl = '//myhost/channel.html'; // Init the SDK upon load window.fbAsyncInit = function() { FB.init({ appId : appID, // App ID channelUrl : channelUrl, status : true, // check login status cookie : true, // enable cookies to allow the server to access the session xfbml : true // parse XFBML }); FB.Event.subscribe('auth.statusChange', function(response) { if(!response.authResponse) FB.login(finish, {scope: 'publish_actions,publish_stream'}); else finish(response); }); FB.getLoginStatus(finish); } function finish(response) { alert("Hello "+response.name); } //]]> </script> </head> <body> <h1>Facebook login</h1> <p>Do NOT close this window.</p> <p>please wait...</p> </body> </html>

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  • [C#][XNA] Draw() 20,000 32 by 32 Textures or 1 Large Texture 20,000 Times

    - by Rudi
    The title may be confusing - sorry about that, it's a poor summary. Here's my dilemma. I'm programming in C# using the .NET Framework 4, and aiming to make a tile-based game with XNA. I have one large texture (256 pixels by 4096 pixels). Remember this is a tile-based game, so this texture is so massive only because it contains many tiles, which are each 32 pixels by 32 pixels. I think the experts will definitely know what a tile-based game is like. The orientation is orthogonal (like a chess board), not isometric. In the Game.Draw() method, I have two choices, one of which will be incredibly more efficient than the other. Choice/Method #1: Semi-Pseudocode: public void Draw() { // map tiles are drawn left-to-right, top-to-bottom for (int x = 0; x < mapWidth; x++) { for (int y = 0; y < mapHeight; y++) { SpriteBatch.Draw( MyLargeTexture, // One large 256 x 4096 texture new Rectangle(x, y, 32, 32), // Destination rectangle - ignore this, its ok new Rectangle(x, y, 32, 32), // Notice the source rectangle 'cuts out' 32 by 32 squares from the texture corresponding to the loop Color.White); // No tint - ignore this, its ok } } } Caption: So, effectively, the first method is referencing one large texture many many times, each time using a small rectangle of this large texture to draw the appropriate tile image. Choice/Method #2: Semi-Pseudocode: public void Draw() { // map tiles are drawn left-to-right, top-to-bottom for (int x = 0; x < mapWidth; x++) { for (int y = 0; y < mapHeight; y++) { Texture2D tileTexture = map.GetTileTexture(x, y); // Getting a small 32 by 32 texture (different each iteration of the loop) SpriteBatch.Draw( tileTexture, new Rectangle(x, y, 32, 32), // Destination rectangle - ignore this, its ok new Rectangle(0, 0, tileTexture.Width, tileTexture.Height), // Notice the source rectangle uses the entire texture, because the entire texture IS 32 by 32 Color.White); // No tint - ignore this, its ok } } } Caption: So, effectively, the second method is drawing many small textures many times. The Question: Which method and why? Personally, I would think it would be incredibly more efficient to use the first method. If you think about what that means for the tile array in a map (think of a large map with 2000 by 2000 tiles, let's say), each Tile object would only have to contain 2 integers, for the X and Y positions of the source rectangle in the one large texture - 8 bytes. If you use method #2, however, each Tile object in the tile array of the map would have to store a 32by32 Texture - an image - which has to allocate memory for the R G B A pixels 32 by 32 times - is that 4096 bytes per tile then? So, which method and why? First priority is speed, then memory-load, then efficiency or whatever you experts believe.

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  • Accessing a vector<vector<int>> as a flat array

    - by user1762276
    For this array: vector<vector<int> > v; v.push_back(vector<int>(0)); v.back().push_back(1); v.back().push_back(2); v.back().push_back(3); v.back().push_back(4); I can output {1, 2, 3, 4} easily enough: cout << v[0][0] << endl; cout << v[0][1] << endl; cout << v[0][2] << endl; cout << v[0][3] << endl; To access it as a flat array I can do this: int* z = (int*)&v[0].front(); cout << z[0] << endl; cout << z[1] << endl; cout << z[2] << endl; cout << z[3] << endl; Now, how do I access the multidimensional vector as a flat multidimensional array? I cannot use the same format as accessing a single-dimensional vector: // This does not work (outputs garbage) int** n = (int**)&v.front(); cout << n[0][0] << endl; cout << n[0][1] << endl; cout << n[0][2] << endl; cout << n[0][3] << endl; The workaround I've found is to do this: int** n = new int* [v.size()]; for (size_t i = 0; i < v.size(); i++) { n[i] = &v.at(i).front(); } cout << n[0][0] << endl; cout << n[0][1] << endl; cout << n[0][2] << endl; cout << n[0][3] << endl; Is there a way to access the entire multidimensional vector like a flat c-style array without having to dynamically allocate each dimension above the data before accessing it? Speed is not critical in the implementation and clarity for maintenance is paramount. A multidimensional vector is just fine for storing the data. However, I want to also expose the data as a flat c-style array in the SDK so that it can be easily accessible by other languages. This means that exposing the vectors as an STL object is a no go. The solution I came up with works fine for my needs as I only evaluate the array once at the very end of processing to "flatten" it. However, is there a better way to go about this? Or am I already doing it the best way I possibly can without re-implementing my own data structure (overkill since my flatten code is only a few lines). Thank you for your advice, friends!

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  • How can I get the previous logged events when a particular logger is triggered?

    - by Ben Laan
    I need to show the previous 10 events when a particular logger is triggered. The goal is to show what previous steps occurred immediately before NHibernate.SQL logging was issued. Currently, I am logging NHibernate sql to a separate file - this is working correctly. <appender name="NHibernateSqlAppender" type="log4net.Appender.RollingFileAppender"> <file value="Logs\NHibernate.log" /> <appendToFile value="true" /> <rollingStyle value="Size" /> <maxSizeRollBackups value="10" /> <maximumFileSize value="10000KB" /> <staticLogFileName value="true" /> <layout type="log4net.Layout.PatternLayout"> <conversionPattern value="%d{dd/MM/yy HH:mm:ss,fff} [%t] %-5p %c - %m%n" /> </layout> </appender> <logger name="NHibernate.SQL" additivity="false"> <level value="ALL"/> <appender-ref ref="NHibernateSqlAppender"/> </logger> <logger name="NHibernate" additivity="false"> <level value="WARN"/> <appender-ref ref="NHibernateSqlAppender"/> </logger> But this only outputs SQL, without context. I would like all previous logs within a specified namespace to also be logged, but only when the HNibernate.SQL appender is triggered. I have investigated the use of BufferingForwardingAppender as a means to collect all events, and then filter them within the NHibernateSqlAppender, but this is not working. I have read about the LoggerMatchFilter class, which seems like it is going to help, but I'm not sure where to put it. <appender name="BufferingForwardingAppender" type="log4net.Appender.BufferingForwardingAppender" > <bufferSize value="10" /> <lossy value="true" /> <evaluator type="log4net.Core.LevelEvaluator"> <threshold value="ALL"/> </evaluator> <appender-ref ref="NHibernateSqlAppender" /> </appender> <appender name="NHibernateSqlAppender" type="log4net.Appender.RollingFileAppender"> <file value="Logs\NHibernate.log" /> <appendToFile value="true" /> <rollingStyle value="Size" /> <maxSizeRollBackups value="10" /> <maximumFileSize value="10000KB" /> <staticLogFileName value="true" /> <filter type="log4net.Filter.LoggerMatchFilter"> <loggerToMatch value="NHibernate.SQL" /> <loggerToMatch value="Laan" /> </filter> <filter type="log4net.Filter.LoggerMatchFilter"> <loggerToMatch value="NHibernate" /> <acceptOnMatch value="false"/> </filter> <layout type="log4net.Layout.PatternLayout"> <conversionPattern value="%d{dd/MM/yy HH:mm:ss,fff} [%t] %-5p %c - %m%n" /> </layout> </appender> <root> <level value="ALL" /> <appender-ref ref="BufferingForwardingAppender"/> </root> The idea is that buffering appender will store all events, but then the NHibernateSqlAppender will only flush when an NHibernate.SQL event fires, plus it will flush the buffer (of 10 previous items, within the specified logger level, which in this example is Laan.*).

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  • Survey Data Model - How to avoid EAV and excessive denormalization?

    - by AlexDPC
    Hi everyone, My database skills are mediocre at best and I have to design a data model for survey data. I have spent some thoughts on this and right now I feel that I am stuck between some kind of EAV model and a design involving hundreds of tables, each with hundreds of columns (and thousands of records). There must be a better way to do this and I hope that the wise folks on this forum can help me. I have already searched various forums, but I couldn't really find a solution. If it has already been given elsewhere, please excuse me and provide me with a link so I can read it up. Some assumptions about the data I have to deal with: Each survey consists of 1 to n questionnaires Each questionnaire consists of 100-2,000 questions (please ignore that 2,000 questions really sound like a lot to answer...) Questions can be of various types: multiple-choice, free text, a number (like age, income, percentages, ...) Each survey involves 10-200 countries (These are not the respondents. The respondents are actually people in the countries.) Depending on the type of questionnaire, each questionnaire is answered by 100-20,000 respondents per country. A country can adapt the questionnaires for a survey, i.e. add, remove or edit questions The data for one country is gathered in a separate database in that country. There is no possibility for online integration from the start. The data for all countries has to be integrated later. This means for example, if a country has deleted a question, that data must somehow be derived from what they sent in order to achieve a uniform design across all countries I will have to write the integration and cleaning software, which will need to work with every country's data In the end the data needs to be exported to flat files, one rectangular grid per country and questionnaire. I have already discussed this topic with people from various backgrounds and have not come to a good solution yet. I mainly got two kinds of opinions. The domain experts, who are used to working with flat files (spreadsheet-style) for data processing and analysis vote for a denormalized structure with loads of tables and columns as I described above (1 table per country and questionnaire). This sounds terrible to me, because I learned that wide tables are to be avoided, it will be annoying to determine which columns are actually in a table when working with it, the database will become cluttered with hundreds of tables (or I even need to set up multiple databases, each with a similar yet a bit differetn design), etc. O-O-programmers vote for a strongly "normalized" design, which would effectively lead to a central table containing all the answers from all respondents to all questions. This table would either need to contain a column of type sql_variant type or multiple answer columns with different types to store answers of different types (multiple choice, free text, ..). The former would essentially be a EAV model. I tend to follow Joe Celko here, who strongly discourages its use (he calls it OTLT or "One True Lookup Table"). The latter would imply that each row would contain null cells for the not applicable types by design. Another alternative I could think of would be to create one table per answer type, i.e., one for multiple-choice questions, one for free text questions, etc.. That's not so generic, it would lead to a lot of union joins, I think and I would have to add a table if a new answer type is invented. Sorry for boring you with all this text and thank you for your input! Cheers, Alex PS: I asked the same question here: http://www.eggheadcafe.com/community/aspnet/13/10242616/survey-data-model--how-to-avoid-eav-and-excessive-denormalization.aspx

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