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  • To Interface or Not?: Creating a polymorphic model relationship in Ruby on Rails dynamically..

    - by Globalkeith
    Please bear with me for a moment as I try to explain exactly what I would like to achieve. In my Ruby on Rails application I have a model called Page. It represents a web page. I would like to enable the user to arbitrarily attach components to the page. Some examples of "components" would be Picture, PictureCollection, Video, VideoCollection, Background, Audio, Form, Comments. Currently I have a direct relationship between Page and Picture like this: class Page < ActiveRecord::Base has_many :pictures, :as => :imageable, :dependent => :destroy end class Picture < ActiveRecord::Base belongs_to :imageable, :polymorphic => true end This relationship enables the user to associate an arbitrary number of Pictures to the page. Now if I want to provide multiple collections i would need an additional model: class PictureCollection < ActiveRecord::Base belongs_to :collectionable, :polymorphic => true has_many :pictures, :as => :imageable, :dependent => :destroy end And alter Page to reference the new model: class Page < ActiveRecord::Base has_many :picture_collections, :as => :collectionable, :dependent => :destroy end Now it would be possible for the user to add any number of image collections to the page. However this is still very static in term of the :picture_collections reference in the Page model. If I add another "component", for example :video_collections, I would need to declare another reference in page for that component type. So my question is this: Do I need to add a new reference for each component type, or is there some other way? In Actionscript/Java I would declare an interface Component and make all components implement that interface, then I could just have a single attribute :components which contains all of the dynamically associated model objects. This is Rails, and I'm sure there is a great way to achieve this, but its a tricky one to Google. Perhaps you good people have some wise suggestions. Thanks in advance for taking the time to read and answer this.

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  • Template function as a template argument

    - by Kos
    I've just got confused how to implement something in a generic way in C++. It's a bit convoluted, so let me explain step by step. Consider such code: void a(int) { // do something } void b(int) { // something else } void function1() { a(123); a(456); } void function2() { b(123); b(456); } void test() { function1(); function2(); } It's easily noticable that function1 and function2 do the same, with the only different part being the internal function. Therefore, I want to make function generic to avoid code redundancy. I can do it using function pointers or templates. Let me choose the latter for now. My thinking is that it's better since the compiler will surely be able to inline the functions - am I correct? Can compilers still inline the calls if they are made via function pointers? This is a side-question. OK, back to the original point... A solution with templates: void a(int) { // do something } void b(int) { // something else } template<void (*param)(int) > void function() { param(123); param(456); } void test() { function<a>(); function<b>(); } All OK. But I'm running into a problem: Can I still do that if a and b are generics themselves? template<typename T> void a(T t) { // do something } template<typename T> void b(T t) { // something else } template< ...param... > // ??? void function() { param<SomeType>(someobj); param<AnotherType>(someotherobj); } void test() { function<a>(); function<b>(); } I know that a template parameter can be one of: a type, a template type, a value of a type. None of those seems to cover my situation. My main question is hence: How do I solve that, i.e. define function() in the last example? (Yes, function pointers seem to be a workaround in this exact case - provided they can also be inlined - but I'm looking for a general solution for this class of problems).

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  • Releasing the keyboard stops shake events. Why?

    - by Moshe
    1) How do I make a UITextField resign the keyboard and hide it? The keyboard is in a dynamically created subview whose superview looks for shake events. Resigning first responder seems to break the shake event handler. 2) how do you make the view holding the keyboard transparent, like see through glass? I have seen this done before. This part has been taken care of thanks guys. As always, code samples are appreciated. I've added my own to help explain the problem. EDIT: Basically, - (void)motionBegan:(UIEventSubtype)motion withEvent:(UIEvent *)event; gets called in my main view controller to handle shaking. When a user taps on the "edit" icon (a pen, in the bottom of the screen - not the traditional UINavigationBar edit button), the main view adds a subview to itself and animates it on to the screen using a custom animation. This subview contains a UINavigationController which holds a UITableView. The UITableView, when a cell is tapped on, loads a subview into itself. This second subview is the culprit. For some reason, a UITextField in this second subview is causing problems. When a user taps on the view, the main view will not respond to shakes unless the UITextField is active (in editing mode?). Additional info: My Motion Event Handler: - (void)motionBegan:(UIEventSubtype)motion withEvent:(UIEvent *)event { NSLog(@"%@", [event description]); SystemSoundID SoundID; NSString *soundFile = [[NSBundle mainBundle] pathForResource:@"shake" ofType:@"aif"]; AudioServicesCreateSystemSoundID((CFURLRef)[NSURL fileURLWithPath:soundFile], &SoundID); AudioServicesPlayAlertSound(SoundID); [self genRandom:TRUE]; } The genRandom: Method: /* Generate random label and apply it */ -(void)genRandom:(BOOL)deviceWasShaken{ if(deviceWasShaken == TRUE){ decisionText.text = [NSString stringWithFormat: (@"%@", [shakeReplies objectAtIndex:(arc4random() % [shakeReplies count])])]; }else{ SystemSoundID SoundID; NSString *soundFile = [[NSBundle mainBundle] pathForResource:@"string" ofType:@"aif"]; AudioServicesCreateSystemSoundID((CFURLRef)[NSURL fileURLWithPath:soundFile], &SoundID); AudioServicesPlayAlertSound(SoundID); decisionText.text = [NSString stringWithFormat: (@"%@", [pokeReplies objectAtIndex:(arc4random() % [pokeReplies count])])]; } } shakeReplies and pokeReplies are both NSArrays of strings. One is used for when a certain part of the screen is poked and one is for when the device is shaken. The app will randomly choose a string from the NSArray and display onscreen. For those of you who work graphically, here is a diagram of the view hierarchy: Root View -> UINavigationController -> UITableView -> Edit View -> Problem UITextfield

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  • Why do sockets not die when server dies? Why does a socket die when server is alive?

    - by Roman
    I try to play with sockets a bit. For that I wrote very simple "client" and "server" applications. Client: import java.net.*; public class client { public static void main(String[] args) throws Exception { InetAddress localhost = InetAddress.getLocalHost(); System.out.println("before"); Socket clientSideSocket = null; try { clientSideSocket = new Socket(localhost,12345,localhost,54321); } catch (ConnectException e) { System.out.println("Connection Refused"); } System.out.println("after"); if (clientSideSocket != null) { clientSideSocket.close(); } } } Server: import java.net.*; public class server { public static void main(String[] args) throws Exception { ServerSocket listener = new ServerSocket(12345); while (true) { Socket serverSideSocket = listener.accept(); System.out.println("A client-request is accepted."); } } } And I found a behavior that I cannot explain: I start a server, than I start a client. Connection is successfully established (client stops running and server is running). Then I close the server and start it again in a second. After that I start a client and it writes "Connection Refused". It seems to me that the server "remember" the old connection and does not want to open the second connection twice. But I do not understand how it is possible. Because I killed the previous server and started a new one! I do not start the server immediately after the previous one was killed (I wait like 20 seconds). In this case the server "forget" the socket from the previous server and accepts the request from the client. I start the server and then I start the client. Connection is established (server writes: "A client-request is accepted"). Then I wait a minute and start the client again. And server (which was running the whole time) accept the request again! Why? The server should not accept the request from the same client-IP and client-port but it does!

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  • Is "Systems Designer" the job title that best describes what I do? [closed]

    - by ivo-rossi
    After having worked as Java developer for almost 3 years in the same company that I currently work at, I moved to a new position associated with the development of the same application. I’m in this new position for more than 1 year now. My official job title is Systems Designer, but I’m not sure this is a title that expresses well what I do. So my question here is what would be the most appropriate job title for me? I see this question as important for my career development. After all, I should be able to explain in one word what I do. And it’s no longer “Java Developer”. Well, in more than one word, this is what I do: The business analysts gather requirements / business problems to be solved with the clients and then discuss these requirements with me. Given the requirements, I design the high level solutions to be implemented in our system (e.g. a new screen on the client application, modifications to existing reports, extension to the XML export format of some objects, etc). I base my decision on the current capabilities of the system, the overall impact that the solutions would have on the system and the estimated effort to implement them (as I was a developer of this same application for almost 3 years before I moved to this position, I’m confident in my estimates). The solutions are discussed iteratively with the business analysts until we agree that they are good. The outcome of this analysis is what we call the “requirements design” document, which is written by me, shared with clients for approval and then also with the team that is going to implement the solutions and test them. Note that there are a few problems that I need to find a solution for that are non-functional. If the users are unhappy with the performance of a certain tool, I will investigate what can be done to speed it up. I will do some research – often based in the Java code itself - to identify possibilities of optimizations. But in this new position I no longer code, the main outcome of my work is really the “requirements design”. Is “Systems Designer” really the most appropriate job title?

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  • Why does WebSharingAppDemo-CEProviderEndToEnd sample still need a client db connection after scope c

    - by Don
    I'm researching a way to build an n-tierd sync solution. From the WebSharingAppDemo-CEProviderEndToEnd sample it seems almost feasable however for some reason, the app will only sync if the client has a live SQL db connection. Can some one explain what I'm missing and how to sync without exposing SQL to the internet? The problem I'm experiencing is that when I provide a Relational sync provider that has an open SQL connection from the client, then it works fine but when I provide a Relational sync provider that has a closed but configured connection string, as in the example, I get an error from the WCF stating that the server did not receive the batch file. So what am I doing wrong? SqlConnectionStringBuilder builder = new SqlConnectionStringBuilder(); builder.DataSource = hostName; builder.IntegratedSecurity = true; builder.InitialCatalog = "mydbname"; builder.ConnectTimeout = 1; provider.Connection = new SqlConnection(builder.ToString()); // provider.Connection.Open(); **** un-commenting this causes the code to work** //create anew scope description and add the appropriate tables to this scope DbSyncScopeDescription scopeDesc = new DbSyncScopeDescription(SyncUtils.ScopeName); //class to be used to provision the scope defined above SqlSyncScopeProvisioning serverConfig = new SqlSyncScopeProvisioning(); .... The error I get occurs in this part of the WCF code: public SyncSessionStatistics ApplyChanges(ConflictResolutionPolicy resolutionPolicy, ChangeBatch sourceChanges, object changeData) { Log("ProcessChangeBatch: {0}", this.peerProvider.Connection.ConnectionString); DbSyncContext dataRetriever = changeData as DbSyncContext; if (dataRetriever != null && dataRetriever.IsDataBatched) { string remotePeerId = dataRetriever.MadeWithKnowledge.ReplicaId.ToString(); //Data is batched. The client should have uploaded this file to us prior to calling ApplyChanges. //So look for it. //The Id would be the DbSyncContext.BatchFileName which is just the batch file name without the complete path string localBatchFileName = null; if (!this.batchIdToFileMapper.TryGetValue(dataRetriever.BatchFileName, out localBatchFileName)) { //Service has not received this file. Throw exception throw new FaultException<WebSyncFaultException>(new WebSyncFaultException("No batch file uploaded for id " + dataRetriever.BatchFileName, null)); } dataRetriever.BatchFileName = localBatchFileName; } Any ideas?

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  • What database table structure should I use for versions, codebases, deployables?

    - by Zac Thompson
    I'm having doubts about my table structure, and I wonder if there is a better approach. I've got a little database for version control repositories (e.g. SVN), the packages (e.g. Linux RPMs) built therefrom, and the versions (e.g. 1.2.3-4) thereof. A given repository might produce no packages, or several, but if there are more than one for a given repository then a particular version for that repository will indicate a single "tag" of the codebase. A particular version "string" might be used to tag a version of the source code in more than one repository, but there may be no relationship between "1.0" for two different repos. So if packages P and Q both come from repo R, then P 1.0 and Q 1.0 are both built from the 1.0 tag of repo R. But if package X comes from repo Y, then X 1.0 has no relationship to P 1.0. In my (simplified) model, I have the following tables (the x_id columns are auto-incrementing surrogate keys; you can pretend I'm using a different primary key if you wish, it's not really important): repository - repository_id - repository_name (unique) ... version - version_id - version_string (unique for a particular repository) - repository_id ... package - package_id - package_name (unique) - repository_id ... This makes it easy for me to see, for example, what are valid versions of a given package: I can join with the version table using the repository_id. However, suppose I would like to add some information to this database, e.g., to indicate which package versions have been approved for release. I certainly need a new table: package_version - version_id - package_id - package_version_released ... Again, the nature of the keys that I use are not really important to my problem, and you can imagine that the data column is "promotion_level" or something if that helps. My doubts arise when I realize that there's really a very close relationship between the version_id and the package_id in my new table ... they must share the same repository_id. Only a small subset of package/version combinations are valid. So I should have some kind of constraint on those columns, enforcing that ... ... I don't know, it just feels off, somehow. Like I'm including somehow more information than I really need? I don't know how to explain my hesitance here. I can't figure out which (if any) normal form I'm violating, but I also can't find an example of a schema with this sort of structure ... not being a DBA by profession I'm not sure where to look. So I'm asking: am I just being overly sensitive?

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  • thread management in nbody code of cuda-sdk

    - by xnov
    When I read the nbody code in Cuda-SDK, I went through some lines in the code and I found that it is a little bit different than their paper in GPUGems3 "Fast N-Body Simulation with CUDA". My questions are: First, why the blockIdx.x is still involved in loading memory from global to share memory as written in the following code? for (int tile = blockIdx.y; tile < numTiles + blockIdx.y; tile++) { sharedPos[threadIdx.x+blockDim.x*threadIdx.y] = multithreadBodies ? positions[WRAP(blockIdx.x + q * tile + threadIdx.y, gridDim.x) * p + threadIdx.x] : //this line positions[WRAP(blockIdx.x + tile, gridDim.x) * p + threadIdx.x]; //this line __syncthreads(); // This is the "tile_calculation" function from the GPUG3 article. acc = gravitation(bodyPos, acc); __syncthreads(); } isn't it supposed to be like this according to paper? I wonder why sharedPos[threadIdx.x+blockDim.x*threadIdx.y] = multithreadBodies ? positions[WRAP(q * tile + threadIdx.y, gridDim.x) * p + threadIdx.x] : positions[WRAP(tile, gridDim.x) * p + threadIdx.x]; Second, in the multiple threads per body why the threadIdx.x is still involved? Isn't it supposed to be a fix value or not involving at all because the sum only due to threadIdx.y if (multithreadBodies) { SX_SUM(threadIdx.x, threadIdx.y).x = acc.x; //this line SX_SUM(threadIdx.x, threadIdx.y).y = acc.y; //this line SX_SUM(threadIdx.x, threadIdx.y).z = acc.z; //this line __syncthreads(); // Save the result in global memory for the integration step if (threadIdx.y == 0) { for (int i = 1; i < blockDim.y; i++) { acc.x += SX_SUM(threadIdx.x,i).x; //this line acc.y += SX_SUM(threadIdx.x,i).y; //this line acc.z += SX_SUM(threadIdx.x,i).z; //this line } } } Can anyone explain this to me? Is it some kind of optimization for faster code?

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  • MySQL LEFT OUTER JOIN virtual table

    - by user1707323
    I am working on a pretty complicated query let me try to explain it to you. Here is the tables that I have in my MySQL database: students Table --- `students` --- student_id first_name last_name current_status status_change_date ------------ ------------ ----------- ---------------- -------------------- 1 John Doe Active NULL 2 Jane Doe Retread 2012-02-01 students_have_courses Table --- `students_have_courses` --- students_student_id courses_course_id s_date e_date int_date --------------------- ------------------- ---------- ---------- ----------- 1 1 2012-01-01 2012-01-04 2012-01-05 1 2 2012-01-05 NULL NULL 2 1 2012-01-10 2012-01-11 NULL students_have_optional_courses Table --- `students_have_optional_courses` --- students_student_id optional_courses_opcourse_id s_date e_date --------------------- ------------------------------ ---------- ---------- 1 1 2012-01-02 2012-01-03 1 1 2012-01-06 NULL 1 5 2012-01-07 NULL Here is my query so far SELECT `students_and_courses`.student_id, `students_and_courses`.first_name, `students_and_courses`.last_name, `students_and_courses`.courses_course_id, `students_and_courses`.s_date, `students_and_courses`.e_date, `students_and_courses`.int_date, `students_have_optional_courses`.optional_courses_opcourse_id, `students_have_optional_courses`.s_date, `students_have_optional_courses`.e_date FROM ( SELECT `c_s_a_s`.student_id, `c_s_a_s`.first_name, `c_s_a_s`.last_name, `c_s_a_s`.courses_course_id, `c_s_a_s`.s_date, `c_s_a_s`.e_date, `c_s_a_s`.int_date FROM ( SELECT `students`.student_id, `students`.first_name, `students`.last_name, `students_have_courses`.courses_course_id, `students_have_courses`.s_date, `students_have_courses`.e_date, `students_have_courses`.int_date FROM `students` LEFT OUTER JOIN `students_have_courses` ON ( `students_have_courses`.`students_student_id` = `students`.`student_id` AND (( `students_have_courses`.`s_date` >= `students`.`status_change_date` AND `students`.current_status = 'Retread' ) OR `students`.current_status = 'Active') ) WHERE `students`.current_status = 'Active' OR `students`.current_status = 'Retread' ) `c_s_a_s` ORDER BY `c_s_a_s`.`courses_course_id` DESC ) `students_and_courses` LEFT OUTER JOIN `students_have_optional_courses` ON ( `students_have_optional_courses`.students_student_id = `students_and_courses`.student_id AND `students_have_optional_courses`.s_date >= `students_and_courses`.s_date AND `students_have_optional_courses`.e_date IS NULL ) GROUP BY `students_and_courses`.student_id; What I want to be returned is the student_id, first_name, and last_name for all Active or Retread students and then LEFT JOIN the highest course_id, s_date, e_date, and int_date for the those students where the s_date is since the status_change_date if status is 'Retread'. Then LEFT JOIN the highest optional_courses_opcourse_id, s_date, and e_date from the students_have_optional_courses TABLE where the students_have_optional_courses.s_date is greater or equal to the students_have_courses.s_date and the students_have_optional_courses.e_date IS NULL Here is what is being returned: student_id first_name last_name courses_course_id s_date e_date int_date optional_courses_opcourse_id s_date_1 e_date_1 ------------ ------------ ----------- ------------------- ---------- ---------- ------------ ------------------------------ ---------- ---------- 1 John Doe 2 2012-01-05 NULL NULL 1 2012-01-06 NULL 2 Jane Doe NULL NULL NULL NULL NULL NULL NULL Here is what I want being returned: student_id first_name last_name courses_course_id s_date e_date int_date optional_courses_opcourse_id s_date_1 e_date_1 ------------ ------------ ----------- ------------------- ---------- ---------- ------------ ------------------------------ ---------- ---------- 1 John Doe 2 2012-01-05 NULL NULL 5 2012-01-07 NULL 2 Jane Doe NULL NULL NULL NULL NULL NULL NULL Everything is working except one thing, I cannot seem to get the highest students_have_optional_courses.optional_courses_opcourse_id no matter how I form the query Sorry, I just solved this myself after writing this all out I think it helped me think of the solution. Here is the solution query: SELECT `students_and_courses`.student_id, `students_and_courses`.first_name, `students_and_courses`.last_name, `students_and_courses`.courses_course_id, `students_and_courses`.s_date, `students_and_courses`.e_date, `students_and_courses`.int_date, `students_optional_courses`.optional_courses_opcourse_id, `students_optional_courses`.s_date, `students_optional_courses`.e_date FROM ( SELECT `c_s_a_s`.student_id, `c_s_a_s`.first_name, `c_s_a_s`.last_name, `c_s_a_s`.courses_course_id, `c_s_a_s`.s_date, `c_s_a_s`.e_date, `c_s_a_s`.int_date FROM ( SELECT `students`.student_id, `students`.first_name, `students`.last_name, `students_have_courses`.courses_course_id, `students_have_courses`.s_date, `students_have_courses`.e_date, `students_have_courses`.int_date FROM `students` LEFT OUTER JOIN `students_have_courses` ON ( `students_have_courses`.`students_student_id` = `students`.`student_id` AND (( `students_have_courses`.`s_date` >= `students`.`status_change_date` AND `students`.current_status = 'Retread' ) OR `students`.current_status = 'Active') ) WHERE `students`.current_status = 'Active' OR `students`.current_status = 'Retread' ) `c_s_a_s` ORDER BY `c_s_a_s`.`courses_course_id` DESC ) `students_and_courses` LEFT OUTER JOIN ( SELECT * FROM `students_have_optional_courses` ORDER BY `students_have_optional_courses`.optional_courses_opcourse_id DESC ) `students_optional_courses` ON ( `students_optional_courses`.students_student_id = `students_and_courses`.student_id AND `students_optional_courses`.s_date >= `students_and_courses`.s_date AND `students_optional_courses`.e_date IS NULL ) GROUP BY `students_and_courses`.student_id;

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  • writing XML with Xerces 3.0.1 and C++ on windows

    - by Jon
    Hi, i have the following function i wrote to create an XML file using Xerces 3.0.1, if i call this function with a filePath of "foo.xml" or "../foo.xml" it works great, but if i pass in "c:/foo.xml" then i get an exception on this line XMLFormatTarget *formatTarget = new LocalFileFormatTarget(targetPath); can someone explain why my code works for relative paths, but not absolute paths please? many thanks. const int ABSOLUTE_PATH_FILENAME_PREFIX_SIZE = 9; void OutputXML(xercesc::DOMDocument* pmyDOMDocument, std::string filePath) { //Return the first registered implementation that has the desired features. In this case, we are after a DOM implementation that has the LS feature... or Load/Save. DOMImplementation *implementation = DOMImplementationRegistry::getDOMImplementation(L"LS"); // Create a DOMLSSerializer which is used to serialize a DOM tree into an XML document. DOMLSSerializer *serializer = ((DOMImplementationLS*)implementation)->createLSSerializer(); // Make the output more human readable by inserting line feeds. if (serializer->getDomConfig()->canSetParameter(XMLUni::fgDOMWRTFormatPrettyPrint, true)) serializer->getDomConfig()->setParameter(XMLUni::fgDOMWRTFormatPrettyPrint, true); // The end-of-line sequence of characters to be used in the XML being written out. serializer->setNewLine(XMLString::transcode("\r\n")); // Convert the path into Xerces compatible XMLCh*. XMLCh *tempFilePath = XMLString::transcode(filePath.c_str()); // Calculate the length of the string. const int pathLen = XMLString::stringLen(tempFilePath); // Allocate memory for a Xerces string sufficent to hold the path. XMLCh *targetPath = (XMLCh*)XMLPlatformUtils::fgMemoryManager->allocate((pathLen + ABSOLUTE_PATH_FILENAME_PREFIX_SIZE) * sizeof(XMLCh)); // Fixes a platform dependent absolute path filename to standard URI form. XMLString::fixURI(tempFilePath, targetPath); // Specify the target for the XML output. XMLFormatTarget *formatTarget = new LocalFileFormatTarget(targetPath); //XMLFormatTarget *myFormTarget = new StdOutFormatTarget(); // Create a new empty output destination object. DOMLSOutput *output = ((DOMImplementationLS*)implementation)->createLSOutput(); // Set the stream to our target. output->setByteStream(formatTarget); // Write the serialized output to the destination. serializer->write(pmyDOMDocument, output); // Cleanup. serializer->release(); XMLString::release(&tempFilePath); delete formatTarget; output->release(); }

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  • devise register confirmation

    - by mattherick
    hello! i have a user and an admin role in my project. i created my authentification with devise, really nice and goot tool for handling the authentification. in my admin role i don´t have any confirmation or something like that. it is really simple and doesn´t make problems. but in my user model i have following things: model: devise :database_authenticatable, :confirmable, :recoverable, :rememberable, :trackable, :validatable, :timeoutable, :registerable # Setup accessible (or protected) attributes for your model attr_accessible :email, :username, :prename, :surname, :phone, :street, :number, :location, :password, :password_confirmation and few validations, but they aren´t relevant this time. my migration looks like following one: class DeviseCreateUsers < ActiveRecord::Migration def self.up create_table(:users) do |t| t.database_authenticatable :null = false t.confirmable t.recoverable t.rememberable t.trackable t.timeoutable t.validateable t.string :username t.string :prename t.string :surname t.string :phone t.string :street t.integer :number t.string :location t.timestamps end add_index :users, :email, :unique => true add_index :users, :confirmation_token, :unique => true add_index :users, :reset_password_token, :unique => true add_index :users, :username, :unique => true add_index :users, :prename, :unique => false add_index :users, :surname, :unique => false add_index :users, :phone, :unique => false add_index :users, :street, :unique => false add_index :users, :number, :unique => false add_index :users, :location, :unique => false end def self.down drop_table :users end end into my route.rb I added following statements: map.devise_for :admins map.devise_for :users, :path_names = { :sign_up = "register", :sign_in = "login" } map.root :controller = "main" and now my problem.. if I register a new user, I fill in all my data in the register form and submit it. After that I get redirected to the controller main with the flash-notice "You have signed up successfully." And I am logged in. But I don´t want to be logged in, because I don´t have confirmed my new user account yet. If I open the console I see the last things in the logs and there I see the confirmation-mail and the text and all stuff, but I am already logged in... I can´t explain why, ... does somebody of you have an idea? If I copy out the confirmation-token from the logs and confirm my account, I can log in, but if I don´t confirm, I also can log in..

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  • Truth tables in code? How to structure state machine?

    - by HanClinto
    I have a (somewhat) large truth table / state machine that I need to implement in my code (embedded C). I anticipate the behavior specification of this state machine to change in the future, and so I'd like to keep this easily modifiable in the future. My truth table has 4 inputs and 4 outputs. I have it all in an Excel spreadsheet, and if I could just paste that into my code with a little formatting, that would be ideal. I was thinking I would like to access my truth table like so: u8 newState[] = decisionTable[input1][input2][input3][input4]; And then I could access the output values with: setOutputPin( LINE_0, newState[0] ); setOutputPin( LINE_1, newState[1] ); setOutputPin( LINE_2, newState[2] ); setOutputPin( LINE_3, newState[3] ); But in order to get that, it looks like I would have to do a fairly confusing table like so: static u8 decisionTable[][][][][] = {{{{ 0, 0, 0, 0 }, { 0, 0, 0, 0 }}, {{ 0, 0, 0, 0 }, { 0, 0, 0, 0 }}}, {{{ 0, 0, 1, 1 }, { 0, 1, 1, 1 }}, {{ 0, 1, 0, 1 }, { 1, 1, 1, 1 }}}}, {{{{ 0, 1, 0, 1 }, { 1, 1, 1, 1 }}, {{ 0, 1, 0, 1 }, { 1, 1, 1, 1 }}}, {{{ 0, 1, 1, 1 }, { 0, 1, 1, 1 }}, {{ 0, 1, 0, 1 }, { 1, 1, 1, 1 }}}}; Those nested brackets can be somewhat confusing -- does anyone have a better idea for how I can keep a pretty looking table in my code? Thanks! Edit based on HUAGHAGUAH's answer: Using an amalgamation of everyone's input (thanks -- I wish I could "accept" 3 or 4 of these answers), I think I'm going to try it as a two dimensional array. I'll index into my array using a small bit-shifting macro: #define SM_INPUTS( in0, in1, in2, in3 ) ((in0 << 0) | (in1 << 1) | (in2 << 2) | (in3 << 3)) And that will let my truth table array look like this: static u8 decisionTable[][] = { { 0, 0, 0, 0 }, { 0, 0, 0, 0 }, { 0, 0, 0, 0 }, { 0, 0, 0, 0 }, { 0, 0, 1, 1 }, { 0, 1, 1, 1 }, { 0, 1, 0, 1 }, { 1, 1, 1, 1 }, { 0, 1, 0, 1 }, { 1, 1, 1, 1 }, { 0, 1, 0, 1 }, { 1, 1, 1, 1 }, { 0, 1, 1, 1 }, { 0, 1, 1, 1 }, { 0, 1, 0, 1 }, { 1, 1, 1, 1 }}; And I can then access my truth table like so: decisionTable[ SM_INPUTS( line1, line2, line3, line4 ) ] I'll give that a shot and see how it works out. I'll also be replacing the 0's and 1's with more helpful #defines that express what each state means, along with /**/ comments that explain the inputs for each line of outputs. Thanks for the help, everyone!

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  • invalid postback event instead of dropdown to datagrid

    - by rima
    I faced with funny situation. I created a page which is having some value, I set these value and control my post back event also. The problem is happening when I change a component index(ex reselect a combobox which is not inside my datagrid) then I dont know why without my page call the Page_Load it goes to create a new row in grid function and all of my parameter are null! I am just receiving null exception. So in other word I try to explain the situation: when I load my page I am initializing some parameter. then everything is working fine. in my page when I change selected item of my combo box, page suppose to go and run function related to that combo box, and call page_load, but it is not going there and it goes to rowcreated function. I am trying to illustrate part of my page. Please help me because I am not receiving any error except null exception and it triger wrong even which seems so complicated for me. public partial class W_CM_FRM_02 : System.Web.UI.Page { protected void Page_Load(object sender, EventArgs e) { if (Page.IsPostBack && !loginFail) return; InitializeItems(); } } private void InitializeItems() { cols = new string[] { "v_classification_code", "v_classification_name" }; arrlstCMMM_CLASSIFICATION = (ArrayList)db.Select(cols, "CMMM_CLASSIFICATION", "v_classification_code <> 'N'", " ORDER BY v_classification_name"); } } protected void DGV_RFA_DETAILS_RowCreated(object sender, GridViewRowEventArgs e) { //db = (Database)Session["oCon"]; foreach (DataRow dr in arrlstCMMM_CLASSIFICATION) ((DropDownList)DGV_RFA_DETAILS.Rows[index].Cells[4].FindControl("OV_RFA_CLASSIFICATION")).Items.Add(new ListItem(dr["v_classification_name"].ToString(), dr["v_classification_code"].ToString())); } protected void V_CUSTOMER_SelectedIndexChanged(object sender, EventArgs e) { if (V_CUSTOMER.SelectedValue == "xxx" || V_CUSTOMER.SelectedValue == "ddd") V_IMPACTED_FUNCTIONS.Enabled = true; } } my form: <%@ Page Language="C#" MasterPageFile="~/MasterPage.master" AutoEventWireup="true" CodeFile="W_CM_FRM_02.aspx.cs" Inherits="W_CM_FRM_02" Title="W_CM_FRM_02" enableeventvalidation="false" EnableViewState="true"%> <td>Project name*</td> <td><asp:DropDownList ID="V_CUSTOMER" runat="server" AutoPostBack="True" onselectedindexchanged="V_CUSTOMER_SelectedIndexChanged" /></td> <td colspan = "8"> <asp:GridView ID="DGV_RFA_DETAILS" runat="server" ShowFooter="True" AutoGenerateColumns="False" CellPadding="1" ForeColor="#333333" GridLines="None" OnRowDeleting="grvRFADetails_RowDeleting" Width="100%" Style="text-align: left" onrowcreated="DGV_RFA_DETAILS_RowCreated"> <RowStyle BackColor="#FFFBD6" ForeColor="#333333" /> <Columns> <asp:BoundField DataField="ON_RowNumber" HeaderText="SNo" /> <asp:TemplateField HeaderText="RFA/RAD/Ticket No*"> <ItemTemplate> <asp:TextBox ID="OV_RFA_NO" runat="server" Width="120"></asp:TextBox> </ItemTemplate> </asp:TemplateField>

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  • MySQL - Calculating fields on the fly vs storing calculated data

    - by Christian Varga
    Hi Everyone, I apologise if this has been asked before, but I can't seem to find an answer to a question that I have about calculating on the fly vs storing fields in a database. I read a few articles that suggested it was preferable to calculate when you can, but I would just like to know if that still applies to the following 2 examples. Example 1. Say you are storing data relating to a car. You store the fuel tank size in litres, and how many litres it uses per 100km. You also want to know how many KMs it can travel, which can be calculated from the tank size and economy. I see 2 ways of doing this: When a car is added or updated, calculate the amount of KMs and store this as a static field in the database. Every time a car is accessed, calculate the amount of KMs on the fly. Because the cars economy/tank size doesn't change (although it could be edited), the KMs is a pretty static value. I don't see why we would calculate it every single time the car is accessed. Wouldn't this waste cpu time as opposed to simply storing it in a separate field in the database and calculating only when a car is added or updated? My next example, which is almost an entirely different question (but on the same topic), relates to counting children. Let's say we have a app which has categories and items. We have a view where we display all the categories, and a count of all the items inside each category. Again, I'm wondering what's better. To perform a MySQL query to count all the items in each category every single time the page is accessed? Or store the count in a field in the categories table and update when an item is added / deleted? I know it is redundant to store anything that can be calculated, but I worry that calculating fields or counting records might be slow as opposed to storing the data in a field. If it's not then please let me know, I just want to learn about when to use either method. On a small scale I guess it wouldn't matter either way, but apps like Facebook, would they really count the amount of friends you have every time someone views your profile or would they just store it as a field? I'd appreciate any responses to both of these scenarios, and any resource that might explain the benefits of calculating vs storing. Thanks in advance, Christian

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  • grdb not working variables

    - by stupid_idiot
    hi, i know this is kinda retarded but I just can't figure it out. I'm debugging this: xor eax,eax mov ah,[var1] mov al,[var2] call addition stop: jmp stop var1: db 5 var2: db 6 addition: add ah,al ret the numbers that I find on addresses var1 and var2 are 0x0E and 0x07. I know it's not segmented, but that ain't reason for it to do such escapades, because the addition call works just fine. Could you please explain to me where is my mistake? I see the problem, dunno how to fix it yet though. The thing is, for some reason the instruction pointer starts at 0x100 and all the segment registers at 0x1628. To address the instruction the used combination is i guess [cs:ip] (one of the segment registers and the instruction pointer for sure). The offset to var1 is 0x10 (probably because from the begining of the code it's the 0x10th byte in order), i tried to examine the memory and what i got was: 1628:100 8 bytes 1628:108 8 bytes 1628:110 <- wtf? (assume another 8 bytes) 1628:118 ... whatever tricks are there in the memory [cs:var1] points somewhere else than in my code, which is probably where the label .data would usually address ds.... probably.. i don't know what is supposed to be at 1628:10 ok, i found out what caused the assness and wasted me whole fuckin day. the behaviour described above is just correct, the code is fully functional. what i didn't know is that grdb debugger for some reason sets the begining address to 0x100... the sollution is to insert the directive ORG 0x100 on the first line and that's the whole thing. the code was working because instruction pointer has the right address to first instruction and goes one by one, but your assembler doesn't know what effective address will be your program stored at so it pretty much remains relative to first line of the code which means all the variables (if not using label for data section) will remain pointing as if it started at 0x0. which of course wouldn't work with DOS. and grdb apparently emulates some DOS features... sry for the language, thx everyone for effort, hope this will spare someone's time if having the same problem... heheh.. at least now i know the reason why to use .data section :))))

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  • Add jquery link to returned text...

    - by Jerry
    Hi all I am trying to add two jquery plugins files to my application. When a user triggers my ajax event, the server will return text with a form button. The plugins (a jquery calendar) will work when the user clicks the form button inside the returned text . I believe I have to add the link inside the return text instead of the main page to let the code work, but not sure how to do this. I am giving out my code and need you experts opinions. Thanks. My main page html //required jquery plugins ...didn't work if I add them in the main application. <script type="text/javascript" src="JS/date.js"></script> <script type="text/javascript" src="JS/datePicker.js"></script> <script type="text/javascript" src="JS/selectWeek.js"></script> <div id="gameInfo"> //return text will be displayed here. </div> My returned text ...part of it.... <form> <div id=returnDiv> // the form input will be added here when a user clicks #addMatch button... </div> <tr> <td><input type="button" id="addMatch" name="addMatch" value="Add Match"/> </td> </tr> </form> My jquery $("#addMatch").live('click', function(){ //the code below will create a calender when a user click the link...I am not sure //where I should add my two jquery plugins link... $("#returnDiv").html("<td><input type='text' size='6' class='date-pick dp-applied'"+ "name='date'><a style='color:white;' class='dp-choose-date' title='Choose Date'"+ "href='#'>Date</a></td>"; return false; }); I hope I explain my question well. +1 to any reply...:D

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  • Android Notification with AlarmManager, Broadcast and Service

    - by user2435829
    this is my code for menage a single notification: myActivity.java public class myActivity extends Activity { protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.mylayout); cal = Calendar.getInstance(); // it is set to 10.30 cal.set(Calendar.HOUR, 10); cal.set(Calendar.MINUTE, 30); cal.set(Calendar.SECOND, 0); long start = cal.getTimeInMillis(); if(cal.before(Calendar.getInstance())) { start += AlarmManager.INTERVAL_FIFTEEN_MINUTES; } Intent mainIntent = new Intent(this, myReceiver.class); pIntent = PendingIntent.getBroadcast(this, 0, mainIntent, PendingIntent.FLAG_UPDATE_CURRENT); AlarmManager myAlarm = (AlarmManager)getSystemService(ALARM_SERVICE); myAlarm.setRepeating(AlarmManager.RTC_WAKEUP, start, AlarmManager.INTERVAL_FIFTEEN_MINUTES, pIntent); } } myReceiver.java public class myReceiver extends BroadcastReceiver { @Override public void onReceive(Context c, Intent i) { Intent myService1 = new Intent(c, myAlarmService.class); c.startService(myService1); } } myAlarmService.java public class myAlarmService extends Service { @Override public IBinder onBind(Intent arg0) { return null; } @Override public void onCreate() { super.onCreate(); } @SuppressWarnings("deprecation") @Override public void onStart(Intent intent, int startId) { super.onStart(intent, startId); displayNotification(); } @Override public void onDestroy() { super.onDestroy(); } public void displayNotification() { Intent mainIntent = new Intent(this, myActivity.class); PendingIntent pIntent = PendingIntent.getActivity(this, 0, mainIntent, PendingIntent.FLAG_UPDATE_CURRENT); NotificationManager nm = (NotificationManager) this.getSystemService(Context.NOTIFICATION_SERVICE); Notification.Builder builder = new Notification.Builder(this); builder.setContentIntent(pIntent) .setAutoCancel(true) .setSmallIcon(R.drawable.ic_noti) .setTicker(getString(R.string.notifmsg)) .setContentTitle(getString(R.string.app_name)) .setContentText(getString(R.string.notifmsg)); nm.notify(0, builder.build()); } } AndroidManifest.xml <uses-permission android:name="android.permission.WAKE_LOCK" /> ... ... ... <service android:name=".myAlarmService" android:enabled="true" /> <receiver android:name=".myReceiver"/> IF the time has NOT past yet everything works perfectly. The notification appears when it must appear. BUT if the time HAS past (let's assume it is 10.31 AM) the notification fires every time... when I close and re-open the app, when I click on the notification... it has a really strange behavior. I can't figure out what's wrong in it. Can you help me please (and explain why, if you find a solution), thanks in advance :)

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  • quick java question

    - by j-unit-122
    private static char[] quicksort (char[] array , int left , int right) { if (left < right) { int p = partition(array , left, right); quicksort(array, left, p - 1 ); quicksort(array, p + 1 , right); } for (char i : array) System.out.print(i + ” ”); System.out.println(); return array; } private static int partition(char[] a, int left, int right) { char p = a[left]; int l = left + 1, r = right; while (l < r) { while (l < right && a[l] < p) l++; while (r > left && a[r] >= p) r--; if (l < r) { char temp = a[l]; a[l] = a[r]; a[r] = temp; } } a[left] = a[r]; a[r] = p; return r; } } hi guys just a quick question regarding the above coding, i know that the above coding returns the following B I G C O M P U T E R B C E G I M P U T O R B C E G I M P U T O R B C E G I M P U T O R B C E G I M P U T O R B C E G I M O P T U R B C E G I M O P R T U B C E G I M O P R T U B C E G I M O P R T U B C E G I M O P R T U B C E G I M O P R T U B C E G I M O P R T U B C E G I M O P R T U when the sequence BIGCOMPUTER is used but my question is can someone explain to me what is happening in the code and how? i know abit about the quick-sort algorithm but it doesnt seem to be the same in the above example.

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  • How do I 'globally' catch exceptions thrown in object instances.

    - by SleepyBobos
    I am currently writing a winforms application (C#). I am making use of the Enterprise Library Exception Handling Block, following a fairly standard approach from what I can see. IE : In the Main method of Program.cs I have wired up event handler to Application.ThreadException event etc. This approach works well and handles the applications exceptional circumstances. In one of my business objects I throw various exceptions in the Set accessor of one of the objects properties set { if (value > MaximumTrim) throw new CustomExceptions.InvalidTrimValue("The value of the minimum trim..."); if (!availableSubMasterWidthSatisfiesAllPatterns(value)) throw new CustomExceptions.InvalidTrimValue("Another message..."); _minimumTrim = value; } My logic for this approach (without turning this into a 'when to throw exceptions' discussion) is simply that the business objects are responsible for checking business rule constraints and throwing an exception that can bubble up and be caught as required. It should be noted that in the UI of my application I do explictly check the values that the public property is being set to (and take action there displaying friendly dialog etc) but with throwing the exception I am also covering the situation where my business object may not be used by a UI eg : the Property is being set by another business object for example. Anyway I think you all get the idea. My issue is that these exceptions are not being caught by the handler wired up to Application.ThreadException and I don't understand why. From other reading I have done the Application.ThreadException event and it handler "... catches any exception that occurs on the main GUI thread". Are the exceptions being raised in my business object not in this thread? I have not created any new threads. I can get the approach to work if I update the code as follows, explicity calling the event handler that is wired to Application.ThreadException. This is the approach outlined in Enterprise Library samples. However this approach requires me to wrap any exceptions thrown in a try catch, something I was trying to avoid by using a 'global' handler to start with. try { if (value > MaximumTrim) throw new CustomExceptions.InvalidTrimValue("The value of the minimum..."); if (!availableSubMasterWidthSatisfiesAllPatterns(value)) throw new CustomExceptions.InvalidTrimValue("Another message"); _minimumTrim = value; } catch (Exception ex) { Program.ThreadExceptionHandler.ProcessUnhandledException(ex); } I have also investigated using wiring a handler up to AppDomain.UnhandledException event but this does not catch the exceptions either. I would be good if someone could explain to me why my exceptions are not being caught by my global exception handler in the first code sample. Is there another approach I am missing or am I stuck with wrapping code in try catch, shown above, as required?

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  • Learning... anything really

    - by WebDevHobo
    I'm particularly interested in Windows PowerShell, but here's a somewhat more general complaint: When asking for help on learning something new, be it a small subject on PHP or understanding a class in Java, what usually happens is that people direct me towards the documentation pages. What I'm looking for is somewhat of a course. A deep explanation of why something works the way it does. I know my basic programming, like Java and C#. I've never seen C or C++, though I have seen a bit of assembler. I know what the Stack and Heap are, how boxing and unboxing works, why you have to deep-copy an array instead of copying the pointer and some other things. Windows PowerShell on the other hand, I know nothing about. And I notice that when reading the small document or some code, I usually forget what it does or why it works. What I am looking for is preferably, a nice tutorial that explains the beginnings, the concepts, and goes to more difficult things at a steady pace. The only thing documentation can do is explain what a function does. That's no good to me since I don't know what I want to do yet. I could read about a thousand functions, and forget about most of them, because I don't need to implement them right after it. Randomly wandering through the documentation doesn't do me any good. So conclude, what is a good tutorial on Windows Powershell? One which explains in clear language what is happening, one which builds on previous things learned. I don't think googling this is a good idea. Doing a Google search on this would turn up numerous tutorials. And experience tells me that you have to look long and hard to find the gem you're looking for. That's why I'm asking here. Because this is the place where you can find more experienced people. Many of the PowerShell guys among you will know the good ones already, and by asking you, I avoid wasting time that could be spent learning. So to summarize: I will not google this!

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  • How to infer the type of a derived class in base class?

    - by enzi
    I want to create a method that allows me to change arbitrary properties of classes that derive from my base class, the result should look like this: SetPropertyValue("size.height", 50); – where size is a property of my derived class and height is a property of size. I'm almost done with my implementation but there's one final obstacle that I want to solve before moving on, to describe this I will first have to explain my implementation a bit: Properties that can be modified are decorated with an attribute There's a method in my base class that searches for all derived classes and their decorated properties For each property I generate a "property modifier", a class that contains 2 delegates: one to set and one to get the value of the property. Property Modifiers are stored in a dictionary, with the name of the property as key In my base class, there is another dictionary that contains all property-modifier-dictionaries, with the Type of the respective class as key. What the SetPropertyValue method does is this: Get the correct property-modifier-dictionary, using the concrete type of the derived class (<- yet to solve) Get the property modifier of the property to change (e.g. of the property size) Use the get or set delegate to modify the property's value Some example code to clarify further: private static Dictionary<RuntimeTypeHandle, object> EditableTypes; //property-modifier-dictionary protected void SetPropertyValue<T>(EditablePropertyMap<T> map, string property, object value) { var property = map[property]; // get the property modifier property.Set((T)this, value); // use the set delegate (encapsulated in a method) } In the above code, T is the Type of the actual (derived) class. I need this type for the get/set delegates. The problem is how to get the EditablePropertyMap<T> when I don't know what T is. My current (ugly) solution is to pass the map in an overriden virtual method in the derived class: public override void SetPropertyValue(string property, object value) { base.SetPropertyValue((EditablePropertyMap<ExampleType>)EditableTypes[typeof(ExampleType)], property, value); } What this does is: get the correct dictionary containing the property modifiers of this class using the class's type, cast it to the appropiate type and pass it to the SetPropertyValue method. I want to get rid of the SetPropertyValue method in my derived class (since there are a lot of derived classes), but don't know yet how to accomplish that. I cannot just make a virtual GetEditablePropertyMap<T> method because I cannot infer a concrete type for T then. I also cannot acces my dictionary directly with a type and retrieve an EditablePropertyMap<T> from it because I cannot cast to it from object in the base class, since again I do not know T. I found some neat tricks to infere types (e.g. by adding a dummy T parameter), but cannot apply them to my specific problem. I'd highly appreciate any suggestions you may have for me.

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  • Abnormally disconnected TCP sockets and write timeout

    - by James
    Hello I will try to explain the problem in shortest possible words. I am using c++ builder 2010. I am using TIdTCPServer and sending voice packets to a list of connected clients. Everything works ok untill any client is disconnected abnormally, For example power failure etc. I can reproduce similar disconnect by cutting the ethernet connection of a connected client. So now we have a disconnected socket but as you know it is not yet detected at server side so server will continue to try to send data to that client too. But when server try to write data to that disconnected client ...... Write() or WriteLn() HANGS there in trying to write, It is like it is wating for somekind of Write timeout. This hangs the hole packet distribution process as a result creating a lag in data transmission to all other clients. After few seconds "Socket Connection Closed" Exception is raised and data flow continues. Here is the code try { EnterCriticalSection(&SlotListenersCriticalSection); for(int i=0;i<SlotListeners->Count;i++) { try { //Here the process will HANG for several seconds on a disconnected socket ((TIdContext*) SlotListeners->Objects[i])->Connection->IOHandler->WriteLn("Some DATA"); }catch(Exception &e) { SlotListeners->Delete(i); } } }__finally { LeaveCriticalSection(&SlotListenersCriticalSection); } Ok i already have a keep alive mechanism which disconnect the socket after n seconds of inactivity. But as you can imagine, still this mechnism cant sync exactly with this braodcasting loop because this braodcasting loop is running almost all the time. So is there any Write timeouts i can specify may be through iohandler or something ? I have seen many many threads about "Detecting disconnected tcp socket" but my problem is little different, i need to avoid that hangup for few seconds during the write attempt. So is there any solution ? Or should i consider using some different mechanism for such data broadcasting for example the broadcasting loop put the data packet in some kind of FIFO buffer and client threads continuously check for available data and pick and deliver it to themselves ? This way if one thread hangs it will not stop/delay the over all distribution thread. Any ideas please ? Thanks for your time and help. Regards Jams

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  • vector related memory allocation question

    - by memC
    hi all, I am encountering the following bug. I have a class Foo . Instances of this class are stored in a std::vector vec of class B. in class Foo, I am creating an instance of class A by allocating memory using new and deleting that object in ~Foo(). the code compiles, but I get a crash at the runtime. If I disable delete my_a from desstructor of class Foo. The code runs fine (but there is going to be a memory leak). Could someone please explain what is going wrong here and suggest a fix? thank you! class A{ public: A(int val); ~A(){}; int val_a; }; A::A(int val){ val_a = val; }; class Foo { public: Foo(); ~Foo(); void createA(); A* my_a; }; Foo::Foo(){ createA(); }; void Foo::createA(){ my_a = new A(20); }; Foo::~Foo(){ delete my_a; }; class B { public: vector<Foo> vec; void createFoo(); B(){}; ~B(){}; }; void B::createFoo(){ vec.push_back(Foo()); }; int main(){ B b; int i =0; for (i = 0; i < 5; i ++){ std::cout<<"\n creating Foo"; b.createFoo(); std::cout<<"\n Foo created"; } std::cout<<"\nDone with Foo creation"; std::cout << "\nPress RETURN to continue..."; std::cin.get(); return 0; }

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  • Benchmark of Java Try/Catch Block

    - by hectorg87
    I know that going into a catch block has some significance cost when executing a program, however, I was wondering if entering a try{} block also had any impact so I started looking for an answer in google with many opinions, but no benchmarking at all. Some answers I found were: Java try/catch performance, is it recommended to keep what is inside the try clause to a minimum? Try Catch Performance Java Java try catch blocks However they didn't answer my question with facts, so I decided to try it for myself. Here's what I did. I have a csv file with this format: host;ip;number;date;status;email;uid;name;lastname;promo_code; where everything after status is optional and will not even have the corresponding ; , so when parsing a validation has to be done to see if the value is there, here's where the try/catch issue came to my mind. The current code that in inherited in my company does this: StringTokenizer st=new StringTokenizer(line,";"); String host = st.nextToken(); String ip = st.nextToken(); String number = st.nextToken(); String date = st.nextToken(); String status = st.nextToken(); String email = ""; try{ email = st.nextToken(); }catch(NoSuchElementException e){ email = ""; } and it repeats what it's done for email with uid, name, lastname and promo_code. and I changed everything to: if(st.hasMoreTokens()){ email = st.nextToken(); } and in fact it performs faster. When parsing a file that doesn't have the optional columns. Here are the average times: --- Trying:122 milliseconds --- Checking:33 milliseconds however, here's what confused me and the reason I'm asking: When running the example with values for the optional columns in all 8000 lines of the CSV, the if() version still performs better than the try/catch version, so my question is Does really the try block does not have any performance impact on my code? The average times for this example are: --- Trying:105 milliseconds --- Checking:43 milliseconds Can somebody explain what's going on here? Thanks a lot

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  • RaphaelJS HTML5 Library pathIntersection() bug or alternative optimisation (screenshots)

    - by user1236048
    I have a chart generated using RaphaelJS library. It is just on long path: M 50 122 L 63.230769230769226 130 L 76.46153846153845 130 L 89.6923076923077 128 L 102.92307692307692 56 L 116.15384615384615 106 L 129.3846153846154 88 L 142.6153846153846 114 L 155.84615384615384 52 L 169.07692307692307 30 L 182.3076923076923 62 L 195.53846153846152 130 L 208.76923076923077 74 L 222 130 L 235.23076923076923 66 L 248.46153846153845 102 L 261.6923076923077 32 L 274.9230769230769 130 L 288.15384615384613 130 L 301.38461538461536 32 L 314.6153846153846 86 L 327.8461538461538 130 L 341.07692307692304 70 L 354.30769230769226 130 L 367.53846153846155 102 L 380.7692307692308 120 L 394 112 L 407.2307692307692 68 L 420.46153846153845 48 L 433.6923076923077 92 L 446.9230769230769 128 L 460.15384615384613 110 L 473.38461538461536 78 L 486.6153846153846 130 L 499.8461538461538 56 L 513.0769230769231 116 L 526.3076923076923 80 L 539.5384615384614 58 L 552.7692307692307 40 L 566 130 L 579.2307692307692 94 L 592.4615384615385 64 L 605.6923076923076 122 L 618.9230769230769 98 L 632.1538461538461 120 L 645.3846153846154 70 L 658.6153846153845 82 L 671.8461538461538 76 L 685.0769230769231 124 L 698.3076923076923 110 L 711.5384615384615 94 L 724.7692307692307 130 L 738 130 L 751.2307692307692 66 L 764.4615384615385 118 L 777.6923076923076 70 L 790.9230769230769 130 L 804.1538461538461 44 L 817.3846153846154 130 L 830.6153846153845 36 L 843.8461538461538 92 L 857.076923076923 130 L 870.3076923076923 76 L 883.5384615384614 130 L 896.7692307692307 60 L 910 88 Also below these chart I have a jqueryUI slider of the same width (860px) and centered with the chart. I want when I move the slider to move a dot on the chart accordingly with the slider position. See attached screenshot: As you can see it seems to work fine. I've implemented this behaviour using the pathIntersection() method. On the slide event at each ui.value (x coordinate) I intersect my chartPath (the one from above) with a vertical straight line at the x coordinate. But still there are some problems. One of them is that it runs very hard, and it kinda freezes sometimes.. and very weird sometimes it doesn't seem to intersect at all even it should.. I'll example below 2 cases I identified: M 499.8461538461538 0 L 499.8461538461538 140 M 910 0 L 910 140 Could you please explain why this intersect behaviour happens (it should return a dot).. and the worst part it seems like it happens randomly.. if I use another chartdata. Also if you can identify another (better) solution to syncronise the slider position with the dot on the chart.. would be perfect. I thought about using Element.getPointAtLength(length), but I don't know how. I think I should save the pathSegments and for each to compute the start Length and the finish Length.

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