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• Database Schema Versioning Strategies

  - by Jack Ryan

  I work on a project that uses a reasonably large database, the live version weighing in at somewhere around 60-80GB. The live database is the only real definitive source of our schema, and because of its size duplicating this database is too slow to be done often. This means we have ended up developing our database schema in a pretty ad hoc way, using sql compare to migrate changes from dev dbs to the live system, and only wiping our dev dbs every month or two. I am hoping to get some pointers on how to improve our database development work flow so that we have a little more control. Some things to think about: Currently nobody is really in charge of the database schema, all developers can change it if they need to, though generally these decisions are talked about before they are done. There are stored procedures, functions, and views in the database. These should probably be dumped to files so they can be reloaded on every build. Schema changes should probably be checked in as scripts. We have started to do this recently. However all our scripts must then be numbered (because there may be dependencies between them), and must be re runnable (because our build script currently runs them all in order). This makes them hard to read because they are full of conditionals that check whether tables or columns already exist. This is a step that is often forgotten by developers. Getting a new database should be quick and easy. This is currently a big problem, it takes several hours to get a copy of last nights backup and restore it onto a dev machine. Some mechanism needs to be in place to allow developers to update static data. We have tables that contain data that is never updated through the application, but does potentially need to be changed when we do a new release (often this drives dropdowns). The whole thing needs to be runnable as part of a build script. Are there any tools that can be used to help to do this? Eventually I would like to be at a point where a new DB can be built from scratch without copying any data from the live system. I don't mind writing some scripts to glue all the steps together but each part should be easily editable so that we continue to use it rather than make changes directly on DBs.

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• How to replace an object in an NSMutableArray at a given index with a new object

  - by shakeelw

  Hi guys. I have an NSMutableArray object(retained, synthesized as all) that is initiated just fine and I can easily add objects to it using the 'addObject:' method. But if I want to replace an object at a certain index with a new one in that NSMutableArray, it doesn't work. For example: ClassA.h @interface ClassA : NSObject { NSMutableArray *list; } @property (nonatomic, copy, readwrite) NSMutableArray *list; end ClassA.m import "ClassA.h" @implementation ClassA @synthesize list; (id)init { [super init]; NSMutableArray *localList = [[NSMutableArray alloc] init]; self.list = localList; [localList release]; //Add initial data [list addObject:@"Hello "]; [list addObject:@"World"]; } // Custom set accessor to ensure the new list is mutable (void)setList:(NSMutableArray *)newList { if (list != newList) { [list release]; list = [newList mutableCopy]; } } -(void)updateTitle:(NSString *)newTitle:(NSString *)theIndex { int i = [theIndex intValue]-1; [self.list replaceObjectAtIndex:i withObject:newTitle]; NSLog((NSString *)[self.list objectAtIndex:i]); // gives the correct output } However, the change remains true only inside the method. from any other method, the NSLog((NSString *)[self.list objectAtIndex:i]); gives the same old value. How can I actually get the old object replaced with the new one at a specific index so that the change can be noticed from within any other method as well. I even modified the method like this, but the result is the same: -(void)updateTitle:(NSString *)newTitle:(NSString *)theIndex { int i = [theIndex intValue]-1; NSMutableArray *localList = [[NSMutableArray alloc] init]; localList = [localList mutableCopy]; for(int j = 0; j < [list count]; j++) { if(j == i) { [localList addObject:newTitle]; NSLog(@"j == 1"); NSLog([NSString stringWithFormat:@"%d", j]); } else { [localList addObject:(NSString *)[self.list objectAtIndex:j]]; } } [self.list release]; //self.list = [localList mutableCopy]; [self setList:localList]; [localList release]; } Please help out guys :)

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• Store Observer not being called always

  - by Nixarn

  Has anyone else here experienced problems with their Store Observer class not being called always when the user for instance cancels a request (or purchases something) We just had our update that brought in app purchases go live last night, and before that we had obviously tested everything tons of times against the Sandbox and everything was working fine. Now however, when the update went live in a real environment we keep getting issues with the store. For instance, in a freshly booted iPhone / iPod, the first time you run the app, if you then try to make a purchase and then immediately cancel it from the first dialog, it seems as if the callback for the cancel is not getting called. If you then restart the app it seems as if it always works after that, or at least. Same thing with other callbacks, seems as if our store observer isn't listening as the callbacks aren't being registered on the phone. One example of this is if you purchase something, then nothing will happen (if this is the first time the app is launched at least). You get the purchase successful dialog from the app store but it seems as if our own code isn't called. If you then quit the app and restart it the callback gets called. Same problem happens if you for instance try to start a request to download all previous purchases and then immediately cancel it as the first dialog pops up, if you do that then the callback for a failed restore is not called, until you then restart the app and try it again, then it always seems to work. The way we have implemented our store observer is by creating a custom class that's implements the SKPaymentTransactionObserver interface. @interface StoreObserver : NSObject<SKPaymentTransactionObserver> In the class we have implemented the following methods: - (void)paymentQueue:(SKPaymentQueue *)queue updatedTransactions:(NSArray *)transactions - (void)paymentQueueRestoreCompletedTransactionsFinished:(SKPaymentQueue *)queue - (void)paymentQueue:(SKPaymentQueue *)queue restoreCompletedTransactionsFailedWithError:(NSError *)error The way our restore process works is that if you tap on the button that allows you to download all we simply run the restoreCompletedTransactions code as follows: [[SKPaymentQueue defaultQueue] restoreCompletedTransactions]; However, the callback, restoreCompletedTransactionsFailedWithError, which has been implemented in the store observer, does not always get called when we try to cancel the request. This happens when you boot the iPhone / iPod and try this for the first time. If you after that restart the app everything works fine. The StoreObserver class is created when our app is launched, just by running the following code: pStoreObserver = [[StoreObserver alloc] init]; [[SKPaymentQueue defaultQueue] addTransactionObserver:pStoreObserver]; Has anyone else had any similar experiences? Or does anyone have any suggestions on how to solve this? As I said, in the sandbox environment everything was working fine, no issues whatsoever, but now once it's gone live we're experiencing these.

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• AuthenticationForm - cookie cross site

  - by bit

  I've 2 web site, the first one myFirst.domain.com and the second one mySecondSite.domain.com. They stay on two different web server and my goal is allow a cross site authentication (my real need is shared authenticationForm Cookie). I've correctly setted web config (machine key node, forms node). The only different is about loginUrl where on myFirstSite appears like "~/login.aspx", instead on mySecondSite it appears like "http://myFirstSite.com/login.aspx". Note that I've not a virtual directory, I've just 2 different web apps. The problem: When I reach myFirstSite login page from mySecondSite I never get redirect from login page, it seems like if cookie doesn't being written. The following is a few of snippet about the issue: MyFirsSite: <machineKey validationKey="..." decryptionKey="..." validation="SHA1" decryption="AES" /> <authentication mode="Forms"> <forms loginUrl="login.aspx" name="authCookie" enableCrossAppRedirects="true"></forms> </authentication> <authorization> <deny users="?" /> <allow users="*"/> </authorization> MyFirstSite code behind: FormsAuthenticationTicket fat = new FormsAuthenticationTicket(1, "userName..", DateTime.Now, DateTime.Now.AddMinutes(30), true, "roles.."); string ticket = FormsAuthentication.Encrypt(fat); HttpCookie authCookie = new HttpCookie(FormsAuthentication.FormsCookieName, ticket); authCookie.Expires = fat.Expiration; authCookie.Domain = "myDomain.com"; Response.Cookies.Add(authCookie); // here other stuff about querystring checking in order to execute exact redirect, however it's not work, I always return on login page MySecondSite: <machineKey validationKey="..." decryptionKey="..." validation="SHA1" decryption="AES"/> <authentication mode="Forms"> <forms loginUrl="http://myFirstSite.domain.com/login.aspx?queryStringToIndicateUrlPage" enableCrossAppRedirects="true"></forms> </authentication> <authorization> Well, that's all. Unfortunately it doesn't works. please, don't pay attention to "queryStringToIndicateUrlPage", it's only simple workaround in order to know whether I must redirect on the same app or on the another one.

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• How can I pipe input to a Java app with Perl?

  - by user319479

  I need to write a Perl script that pipes input into a Java program. This is related to this, but that didn't help me. My issue is that the Java app doesn't get the print statements until I close the handle. What I found online was that $| needs to be set to something greater than 0, in which case newline characters will flush the buffer. This still doesn't work. This is the script: #! /usr/bin/perl -w use strict; use File::Basename; $|=1; open(TP, "| java -jar test.jar") or die "fail"; sleep(2); print TP "this is test 1\n"; print TP "this is test 2\n"; print "tests printed, waiting 5s\n"; sleep(5); print "wait over. closing handle...\n"; close TP; print "closed.\n"; print "sleeping for 5s...\n"; sleep(5); print "script finished!\n"; exit And here is a sample Java app: import java.util.Scanner; public class test{ public static void main( String[] args ){ Scanner sc = new Scanner( System.in ); int crashcount = 0; while( true ){ try{ String input = sc.nextLine(); System.out.println( ":: INPUT: " + input ); if( "bananas".equals(input) ){ break; } } catch( Exception e ){ System.out.println( ":: EXCEPTION: " + e.toString() ); crashcount++; if( crashcount == 5 ){ System.out.println( ":: Looks like stdin is broke" ); break; } } } System.out.println( ":: IT'S OVER!" ); return; } } The Java app should respond to receiving the test prints immediately, but it doesn't until the close statement in the Perl script. What am I doing wrong? Note: the fix can only be in the Perl script. The Java app can't be changed. Also, File::Basename is there because I'm using it in the real script.

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• Keeping Velocity Constant and Player in Position - Sidescrolling

  - by user2904951

  I'm working on a Little Mobile Game with Cocos2D-X and Box2D. The Point where I got stuck is the movement of a box2d-body (the main actor) and the according Sprite. Now I want to : move this Body with a constant velocity along the x-axis, no matter if it's rolling (it's a circleshape) upwards or downwards keep the body nearly sticking to the ground on which it's rolling keep the Body and the according Sprite in the Center of the Screen. What I tried : in the update()- method I used body->SetLinearVelocity(b2Vec2(x,y)) to higher/lower values, if the Body was passing a constant value for his velocity I used to set very high y-Values in body->SetLinearVelocity(b2Vec2(x,y)) First tried to use CCFollow with my playerSprite, which was also Scrolling along the y-axis, as i only need to scroll along the x-axis, so I decided to move the whole layer which is containing the ambience (platforms etc.) to the left of my Screen and my Player Body & Player sprite to the right of the Screen, adjusting the speed values to Keep the Player in the Center of the Screen. Well... ...didn't work as i wanted it to, because each time i set the velocity manually (I also tried to use body->applyLinearImpulse(...) when the Body is moving upwards just as playing around with the value of velocityIterations in world->Step(...)) there's a small delay, which pushes the player Body more or less further of the Center of the Screen. ... didn't also work as I expected it to, because I needed to adjust the x-Values, when the Body was moving upwards to Keep it not getting slowed down, this made my Body even less sticky to the ground.... ... CCFollow did a good Job, except that I didn't want to scroll along the y-axis also and it Forces the overgiven sprite to start in the Center of the Screen. Moving the whole Layer even brought no good results, I have tried a Long time to adjust values of the movement Speed of the layer and the Body to Keep it negating each other, that the player stays nearly in the Center of the Screen.... So my question is : Does anyone of you have any Kind of new Approach for me to solve this cohesive bunch of Problems ? Cheers, Seb

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• Subclassing and adding data members

  - by Marius

  I have an hierarchy of classes that looks like the following: class Critical { public: Critical(int a, int b) : m_a(a), m_b(b) { } virtual ~Critical() { } int GetA() { return m_a; } int GetB() { return m_b; } void SetA(int a) { m_a = a; } void SetB(int b) { m_b = b; } protected: int m_a; int m_b; }; class CriticalFlavor : public Critical { public: CriticalFlavor(int a, int b, int flavor) : Critical(a, b), m_flavor(flavor) { } virtual ~CriticalFlavor() { } int GetFlavor() { return m_flavor; } void SetFlavor(int flavor) { m_flavor = flavor; } protected: int m_flavor; }; class CriticalTwist : public Critical { public: CriticalTwist(int a, int b, int twist) : Critical(a, b), m_twist(twist) { } virtual ~CriticalTwist() { } int GetTwist() { return m_twist; } void SetTwist(int twist) { m_twist = twist; } protected: int m_twist; }; The above does not seem right to me in terms of the design and what bothers me the most is the fact that the addition of member variables seems to drive the interface of these classes (the real code that does the above is a little more complex but still embracing the same pattern). That will proliferate when in need for another "Critical" class that just adds some other property. Does this feel right to you? How could I refactor such code? An idea would be to have just a set of interfaces and use composition when it comes to the base object like the following: class Critical { public: virtual int GetA() = 0; virtual int GetB() = 0; virtual void SetA(int a) = 0; virtual void SetB(int b) = 0; }; class CriticalImpl { public: CriticalImpl(int a, int b) : m_a(a), m_b(b) { } ~CriticalImpl() { } int GetA() { return m_a; } int GetB() { return m_b; } void SetA(int a) { m_a = a; } void SetB(int b) { m_b = b; } private: int m_a; int m_b; }; class CriticalFlavor { public: virtual int GetFlavor() = 0; virtual void SetFlavor(int flavor) = 0; }; class CriticalFlavorImpl : public Critical, public CriticalFlavor { public: CriticalFlavorImpl(int a, int b, int flavor) : m_flavor(flavor), m_critical(new CriticalImpl(a, b)) { } ~CriticalFlavorImpl() { delete m_critical; } int GetFlavor() { return m_flavor; } void SetFlavor(int flavor) { m_flavor = flavor; } int GetA() { return m_critical-GetA(); } int GetB() { return m_critical-GetB(); } void SetA(int a) { m_critical-SetA(a); } void SetB(int b) { m_critical-SetB(b); } private: int m_flavor; CriticalImpl* m_critical; };

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• MySQL search for user and their roles

  - by Jenkz

  I am re-writing the SQL which lets a user search for any other user on our site and also shows their roles. An an example, roles can be "Writer", "Editor", "Publisher". Each role links a User to a Publication. Users can take multiple roles within multiple publications. Example table setup: "users" : user_id, firstname, lastname "publications" : publication_id, name "link_writers" : user_id, publication_id "link_editors" : user_id, publication_id Current psuedo SQL: SELECT * FROM ( (SELECT user_id FROM users WHERE firstname LIKE '%Jenkz%') UNION (SELECT user_id FROM users WHERE lastname LIKE '%Jenkz%') ) AS dt JOIN (ROLES STATEMENT) AS roles ON roles.user_id = dt.user_id At the moment my roles statement is: SELECT dt2.user_id, dt2.publication_id, dt.role FROM ( (SELECT 'writer' AS role, link_writers.user_id, link_writers.publication_id FROM link_writers) UNION (SELECT 'editor' AS role, link_editors.user_id, link_editors.publication_id FROM link_editors) ) AS dt2 The reason for wrapping the roles statement in UNION clauses is that some roles are more complex and require a table join to find the publication_id and user_id. As an example "publishers" might be linked accross two tables "link_publishers": user_id, publisher_group_id "link_publisher_groups": publisher_group_id, publication_id So in that instance, the query forming part of my UNION would be: SELECT 'publisher' AS role, link_publishers.user_id, link_publisher_groups.publication_id FROM link_publishers JOIN link_publisher_groups ON lpg.group_id = lp.group_id I'm pretty confident that my table setup is good (I was warned off the one-table-for-all system when researching the layout). My problem is that there are now 100,000 rows in the users table and upto 70,000 rows in each of the link tables. Initial lookup in the users table is fast, but the joining really slows things down. How can I only join on the relevant roles? -------------------------- EDIT ---------------------------------- Explain above (open in a new window to see full resolution). The bottom bit in red, is the "WHERE firstname LIKE '%Jenkz%'" the third row searches WHERE CONCAT(firstname, ' ', lastname) LIKE '%Jenkz%'. Hence the large row count, but I think this is unavoidable, unless there is a way to put an index accross concatenated fields? The green bit at the top just shows the total rows scanned from the ROLES STATEMENT. You can then see each individual UNION clause (#6 - #12) which all show a large number of rows. Some of the indexes are normal, some are unique. It seems that MySQL isn't optimizing to use the dt.user_id as a comparison for the internal of the UNION statements. Is there any way to force this behaviour? Please note that my real setup is not publications and writers but "webmasters", "players", "teams" etc.

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• How should I ethically approach user password storage for later plaintext retrieval?

  - by Shane

  As I continue to build more and more websites and web applications I am often asked to store user's passwords in a way that they can be retrieved if/when the user has an issue (either to email a forgotten password link, walk them through over the phone, etc.) When I can I fight bitterly against this practice and I do a lot of ‘extra’ programming to make password resets and administrative assistance possible without storing their actual password. When I can’t fight it (or can’t win) then I always encode the password in some way so that it at least isn’t stored as plaintext in the database—though I am aware that if my DB gets hacked that it won’t take much for the culprit to crack the passwords as well—so that makes me uncomfortable. In a perfect world folks would update passwords frequently and not duplicate them across many different sites—unfortunately I know MANY people that have the same work/home/email/bank password, and have even freely given it to me when they need assistance. I don’t want to be the one responsible for their financial demise if my DB security procedures fail for some reason. Morally and ethically I feel responsible for protecting what can be, for some users, their livelihood even if they are treating it with much less respect. I am certain that there are many avenues to approach and arguments to be made for salting hashes and different encoding options, but is there a single ‘best practice’ when you have to store them? In almost all cases I am using PHP and MySQL if that makes any difference in the way I should handle the specifics. Additional Information for Bounty I want to clarify that I know this is not something you want to have to do and that in most cases refusal to do so is best. I am, however, not looking for a lecture on the merits of taking this approach I am looking for the best steps to take if you do take this approach. In a note below I made the point that websites geared largely toward the elderly, mentally challenged, or very young can become confusing for people when they are asked to perform a secure password recovery routine. Though we may find it simple and mundane in those cases some users need the extra assistance of either having a service tech help them into the system or having it emailed/displayed directly to them. In such systems the attrition rate from these demographics could hobble the application if users were not given this level of access assistance, so please answer with such a setup in mind. Thanks to Everyone This has been a fun questions with lots of debate and I have enjoyed it. In the end I selected an answer that both retains password security (I will not have to keep plain text or recoverable passwords), but also makes it possible for the user base I specified to log into a system without the major drawbacks I have found from normal password recovery. As always there were about 5 answers that I would like to have marked correct for different reasons, but I had to choose the best one--all the rest got a +1. Thanks everyone!

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• Move an object in the direction of a bezier curve?

  - by Sent1nel

  I have an object with which I would like to make follow a bezier curve and am a little lost right now as to how to make it do that based on time rather than the points that make up the curve. .::Current System::. Each object in my scene graph is made from position, rotation and scale vectors. These vectors are used to form their corresponding matrices: scale, rotation and translation. Which are then multiplied in that order to form the local transform matrix. A world transform (Usually the identity matrix) is then multiplied against the local matrix transform. class CObject { public: // Local transform functions Matrix4f GetLocalTransform() const; void SetPosition(const Vector3f& pos); void SetRotation(const Vector3f& rot); void SetScale(const Vector3f& scale); // Local transform Matrix4f m_local; Vector3f m_localPostion; Vector3f m_localRotation; // rotation in degrees (xrot, yrot, zrot) Vector3f m_localScale; } Matrix4f CObject::GetLocalTransform() { Matrix4f out(Matrix4f::IDENTITY); Matrix4f scale(), rotation(), translation(); scale.SetScale(m_localScale); rotation.SetRotationDegrees(m_localRotation); translation.SetTranslation(m_localTranslation); out = scale * rotation * translation; } The big question I have are 1) How do I orientate my object to face the tangent of the Bezier curve? 2) How do I move that object along the curve without just setting objects position to that of a point on the bezier cuve? Heres an overview of the function thus far void CNodeControllerPieceWise::AnimateNode(CObject* pSpatial, double deltaTime) { // Get object latest pos. Vector3f posDelta = pSpatial->GetWorldTransform().GetTranslation(); // Get postion on curve Vector3f pos = curve.GetPosition(m_t); // Get tangent of curve Vector3f tangent = curve.GetFirstDerivative(m_t); } Edit: sorry its not very clear. I've been working on this for ages and its making my brain turn to mush. I want the object to be attached to the curve and face the direction of the curve. As for movement, I want to object to follow the curve based on the time this way it creates smooth movement throughout the curve.

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• Get content in iframe to use as much space as it needs

  - by Mark

  I'm trying to write a simple JavaScript based modal dialog. The JavaScript function takes the content, puts it in a new iframe and adds the iframe to the page. Works great so far, the only problem is that the content of the dialog (e.g. a table) gets wrapped, although plenty of space is available on the page. I'd like the content of the dialog, a table in my case, to use as much space as it needs, without wrapping any lines. I tried lots of combinations of setting width/style.width on the iframe and the table. Nothing did the trick. Here the code to show the iframe dialog: function SimpleDialog() { this.domElement = document.createElement('iframe'); this.domElement.setAttribute('style', 'border: 1px solid red; z-index: 201; position: absolute; top: 0px; left: 0px;'); this.showWithContent = function(content) { document.getElementsByTagName('body')[0].appendChild(this.domElement); this.domElement.contentDocument.body.appendChild(content); var contentBody = this.domElement.contentDocument.body; contentBody.style.padding = '0px'; contentBody.style.margin = '0px'; // Set the iframe size to the size of content. // However, content got wrapped already. this.domElement.style.height = content.offsetHeight + 'px'; this.domElement.style.width = content.offsetWidth + 'px'; this._centerOnScreen(); }; this._centerOnScreen = function() { this.domElement.style.left = window.pageXOffset + (window.innerWidth / 2) - (this.domElement.offsetWidth / 2) + 'px'; this.domElement.style.top = window.pageYOffset + (window.innerHeight / 2) - (this.domElement.offsetHeight / 2) + 'px'; }; } Here the test code: var table = document.createElement('table'); table.setAttribute('style', 'border: 1px solid black; width: 100%;'); table.innerHTML = "<tr><td style='font-size:40px;'>Hello world in big letters</td></tr><tr><td>second row</td></tr>"; var dialog = new SimpleDialog(); dialog.showWithContent(table); The table shows up nicely centered on the page, but the words in the first cell are wrapped to two lines. How do I get the table to use as much space as it needs (without using white-space: nowrap ;) Thanks in advance for any suggestions! -Mark

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• Passing custom info to mongrel_rails start

  - by whaka

  One thing I really don't understand is how I can pass custom start-up options to a mongrel instance. I see that a common approach is the use environment variables, but in my environment this is not going to work because my rails application serves many different clients. Much code is shared between clients, but there are also many differences which I implement by subclassing controllers and views to overload or extend existing features or introduce new ones. To make this all work, I simply add the paths to client specific modules the module load path ($:). In order to start the application for a particular client, I could now use an environment variable like say, TARGET=AMAZONE. Unfortunately, on some systems I'm running multiple mongrel clusters, each cluster serving a different client. Some of these systems run under Windows and to start mongrel I installed mongrel_services. Clearly, this makes my environment variable unsuitable. Passing this extra bit of data to the application is proving to be a real challenge. For a start, mongrel_rails service_install will reject any [custom] command line parameters that aren't documented. I'm not too concerned as installing the services using the install program is trivial. However, even if I manage to install mongrel_services such that when run it passes the custom command line option --target to mongrel_rails start, I get an error because mongrel_rails doesn't recognize the switch. So here were the things I looked at: Pass an extra parameter: mongrel_rails start --target XYZ ... use a config file and add target:XYZ, then do: mongrel_rails start -C x:\myapp\myconfig.yml modify the file: Ruby\lib\ruby\gems\1.8\gems\mongrel-1.1.5-x86-mswin32-60\lib\mongrel\command.rb Perhaps I can use the --script option, but all docs that I found on it were for Unix 1 and 2 simply don't work. I played with 4 but never managed it to do anything. So I had no choice but to go with 3. While it is relatively simple, I hate changing ruby library code. Particularly disappointing is that 2 doesn't work. I mean what is so unreasonable about adding other [custom] options in the config file? Actually I think this is a fundamental piece that is missing in rails. Somehow, the application should be able to register and access command line arguments it expects. If anybody has a good idea how to do this more elegantly using the current infrastructure, I have a chocolate fish to give away!!!

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• Big Oh Notation - formal definition.

  - by aloh

  I'm reading a textbook right now for my Java III class. We're reading about Big-Oh and I'm a little confused by its formal definition. Formal Definition: "A function f(n) is of order at most g(n) - that is, f(n) = O(g(n)) - if a positive real number c and positive integer N exist such that f(n) <= c g(n) for all n = N. That is, c g(n) is an upper bound on f(n) when n is sufficiently large." Ok, that makes sense. But hold on, keep reading...the book gave me this example: "In segment 9.14, we said that an algorithm that uses 5n + 3 operations is O(n). We now can show that 5n + 3 = O(n) by using the formal definition of Big Oh. When n = 3, 5n + 3 <= 5n + n = 6n. Thus, if we let f(n) = 5n + 3, g(n) = n, c = 6, N = 3, we have shown that f(n) <= 6 g(n) for n = 3, or 5n + 3 = O(n). That is, if an algorithm requires time directly proportional to 5n + 3, it is O(n)." Ok, this kind of makes sense to me. They're saying that if n = 3 or greater, 5n + 3 takes less time than if n was less than 3 - thus 5n + n = 6n - right? Makes sense, since if n was 2, 5n + 3 = 13 while 6n = 12 but when n is 3 or greater 5n + 3 will always be less than or equal to 6n. Here's where I get confused. They give me another example: Example 2: "Let's show that 4n^2 + 50n - 10 = O(n^2). It is easy to see that: 4n^2 + 50n - 10 <= 4n^2 + 50n for any n. Since 50n <= 50n^2 for n = 50, 4n^2 + 50n - 10 <= 4n^2 + 50n^2 = 54n^2 for n = 50. Thus, with c = 54 and N = 50, we have shown that 4n^2 + 50n - 10 = O(n^2)." This statement doesn't make sense: 50n <= 50n^2 for n = 50. Isn't any n going to make the 50n less than 50n^2? Not just greater than or equal to 50? Why did they even mention that 50n <= 50n^2? What does that have to do with the problem? Also, 4n^2 + 50n - 10 <= 4n^2 + 50n^2 = 54n^2 for n = 50 is going to be true no matter what n is. And how in the world does picking numbers show that f(n) = O(g(n))? Please help me understand! :(

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• jQuery: load refuses to get dynamic content in IE6

  - by user260157

  jQuery refuses to load my dynamic content in IE6. All in FireFox & Safari works fine. Only IE6 is being a pain. When I try the a html with <p>Hello World</p> that works. Properly. But when loading a PHP it doesn't work! As you can see it's doing multiple things. <script type="text/javascript"> // When the document is ready set up our sortable with it's inherant function(s) $(document).ready(function() { // Sort list & amend in database function sortTableMenuAndReload() { var order = $('#menuList').sortable('serialize'); $.post("PLUGINS/SortableMenu/process-sortable.php",order); $("#menuList").load("PLUGINS/SortableMenu/sortableMenu_ajax.php"); } function sortTableOrder() { var order = $('#menuList').sortable('serialize'); $.post("PLUGINS/SortableMenu/process-sortable.php",order); } function sortTableOrderAndRemove(removeID) { $('#listItem_'+removeID).remove(); var order = $('#menuList').sortable('serialize'); $.post("PLUGINS/SortableMenu/process-sortable.php",order); $("#menuList").load("PLUGINS/SortableMenu/sortableMenu_ajax.php"); } $("#menuList > li > .remove").live('click', function () { var removeID = $(this).attr('id'); $.ajax({ type: 'post', url: 'PLUGINS/SortableMenu/removeLine.php', data: 'id='+removeID, success: sortTableOrderAndRemove(removeID) }); }); $("#menuList > li > .publish").live('click', function () { var publishID = $(this).attr('id'); $.ajax({ type: 'post', url: 'PLUGINS/SortableMenu/publishLine.php', data: 'id='+publishID, success: sortTableOrder }); }); $('#new_documents > li').draggable({ addClasses: false, helper:'clone', connectToSortable:'#menuList' }); $("#menuList").droppable({ addClasses: false, drop: function() { var clone = $("#menuList > li#newArticleTYPE1"); $(clone).attr("id","listItem_newArticleTYPE1"); } }); $("#menuList").sortable({ opacity: 0.6, handle : '.handle, .remove', update : sortTableMenuAndReload }); }); </script>

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• PHP readdir(): 3 is not a valid Directory resource

  - by Jordan

  <?php function convert(){ //enable error reporting for debugging error_reporting(E_ALL | E_STRICT); //convert pdf's to html using payroll.sh script and //pdftohtml command line tool $program = "payroll.sh"; $toexec="sh /var/www/html/tmp/" . $program . " 2>&1"; $exec=shell_exec($toexec); //display message from payroll.sh //echo $exec; //echo ('<br/>'); } function process(){ $dir = '/var/www/html/tmp/converted'; //echo ('one'); if (is_dir($dir)) { //echo ('two'); if ($dh = opendir($dir)) { //echo ('three'); while (($file = readdir($dh)) !== false) { //echo ('four'); if ($file != "." && $file != ".."){ echo 'opening file: '; echo $file; echo ("<br/>"); $fp = fopen('/var/www/html/tmp/converted/' . $file, 'r+'); $count = 0; //while file is not at the EOF marker while (!feof($fp)) { $line = fgets($fp); if($count==21) { $employeeID = substr($line,71,4); echo 'employee ID: '; echo $employeeID; echo ('<br/>'); //echo ('six'); $count++; } else if($count==30) { $employeeDate = substr($line,71,10); echo 'employee Date: '; echo $employeeDate; echo ('<br/>'); //echo ('seven'); $count++; } else { //echo ('eight'); //echo ('<br/>'); $count++; } } fclose($fp); closedir($dh); } } } } } convert(); process(); ?> I am setting up a php script that will take a paystub in pdf format, convert it to html, then import it into Drupal after getting the date and employee ID. The code only seems to process the first file in the directory then it gives me this: opening file: dd00000112_28_2010142011-1.html employee ID: 9871 employee Date: 12/31/2010 Warning: readdir(): 3 is not a valid Directory resource in /var/www/html/pay.mistequaygroup.com/payroll.php on line 29 The '3' in the error really confuses me, and google is not helping much. Could it be the 3rd iteration of the loop? The only files in the directory reddir() is scanning are the .html files waiting to be processed. Any ideas? Also, how does my code look? I'm fairly new to doing any real programming and I don't get too much input around work.

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• How to get all captures of subgroup matches with preg_match_all()?

  - by hakre

  Update/Note: I think what I'm probably looking for is to get the captures of a group in PHP. Referenced: PCRE regular expressions using named pattern subroutines. (Read carefully:) I have a string that contains a variable number of segments (simplified): $subject = 'AA BB DD '; // could be 'AA BB DD CC EE ' as well I would like now to match the segments and return them via the matches array: $pattern = '/^(([a-z]+) )+$/i'; $result = preg_match_all($pattern, $subject, $matches); This will only return the last match for the capture group 2: DD. Is there a way that I can retrieve all subpattern captures (AA, BB, DD) with one regex execution? Isn't preg_match_all suitable for this? This question is a generalization. Both the $subject and $pattern are simplified. Naturally with such the general list of AA, BB, .. is much more easy to extract with other functions (e.g. explode) or with a variation of the $pattern. But I'm specifically asking how to return all of the subgroup matches with the preg_...-family of functions. For a real life case imagine you have multiple (nested) level of a variant amount of subpattern matches. Example This is an example in pseudo code to describe a bit of the background. Imagine the following: Regular definitions of tokens: CHARS := [a-z]+ PUNCT := [.,!?] WS := [ ] $subject get's tokenized based on these. The tokenization is stored inside an array of tokens (type, offset, ...). That array is then transformed into a string, containing one character per token: CHARS -> "c" PUNCT -> "p" WS -> "s" So that it's now possible to run regular expressions based on tokens (and not character classes etc.) on the token stream string index. E.g. regex: (cs)?cp to express one or more group of chars followed by a punctuation. As I now can express self-defined tokens as regex, the next step was to build the grammar. This is only an example, this is sort of ABNF style: words = word | (word space)+ word word = CHARS+ space = WS punctuation = PUNCT If I now compile the grammar for words into a (token) regex I would like to have naturally all subgroup matches of each word. words = (CHARS+) | ( (CHARS+) WS )+ (CHARS+) # words resolved to tokens words = (c+)|((c+)s)+c+ # words resolved to regex I could code until this point. Then I ran into the problem that the sub-group matches did only contain their last match. So I have the option to either create an automata for the grammar on my own (which I would like to prevent to keep the grammar expressions generic) or to somewhat make preg_match working for me somehow so I can spare that. That's basically all. Probably now it's understandable why I simplified the question. Related: pcrepattern man page Get repeated matches with preg_match_all()

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• Query performs poorly unless a temp table is used

  - by Paul McLoughlin

  The following query takes about 1 minute to run, and has the following IO statistics: SELECT T.RGN, T.CD, T.FUND_CD, T.TRDT, SUM(T2.UNITS) AS TotalUnits FROM dbo.TRANS AS T JOIN dbo.TRANS AS T2 ON T2.RGN=T.RGN AND T2.CD=T.CD AND T2.FUND_CD=T.FUND_CD AND T2.TRDT<=T.TRDT JOIN TASK_REQUESTS AS T3 ON T3.CD=T.CD AND T3.RGN=T.RGN AND T3.TASK = 'UPDATE_MEM_BAL' GROUP BY T.RGN, T.CD, T.FUND_CD, T.TRDT (4447 row(s) affected) Table 'TRANSACTIONS'. Scan count 5977, logical reads 7527408, physical reads 0, read-ahead reads 0, lob logical reads 0, lob physical reads 0, lob read-ahead reads 0. Table 'TASK_REQUESTS'. Scan count 1, logical reads 11, physical reads 0, read-ahead reads 0, lob logical reads 0, lob physical reads 0, lob read-ahead reads 0. SQL Server Execution Times: CPU time = 58157 ms, elapsed time = 61437 ms. If I instead introduce a temporary table then the query returns quickly and performs less logical reads: CREATE TABLE #MyTable(RGN VARCHAR(20) NOT NULL, CD VARCHAR(20) NOT NULL, PRIMARY KEY([RGN],[CD])); INSERT INTO #MyTable(RGN, CD) SELECT RGN, CD FROM TASK_REQUESTS WHERE TASK='UPDATE_MEM_BAL'; SELECT T.RGN, T.CD, T.FUND_CD, T.TRDT, SUM(T2.UNITS) AS TotalUnits FROM dbo.TRANS AS T JOIN dbo.TRANS AS T2 ON T2.RGN=T.RGN AND T2.CD=T.CD AND T2.FUND_CD=T.FUND_CD AND T2.TRDT<=T.TRDT JOIN #MyTable AS T3 ON T3.CD=T.CD AND T3.RGN=T.RGN GROUP BY T.RGN, T.CD, T.FUND_CD, T.TRDT (4447 row(s) affected) Table 'Worktable'. Scan count 5974, logical reads 382339, physical reads 0, read-ahead reads 0, lob logical reads 0, lob physical reads 0, lob read-ahead reads 0. Table 'TRANSACTIONS'. Scan count 4, logical reads 4547, physical reads 0, read-ahead reads 0, lob logical reads 0, lob physical reads 0, lob read-ahead reads 0. Table '#MyTable________________________________________________________________000000000013'. Scan count 1, logical reads 2, physical reads 0, read-ahead reads 0, lob logical reads 0, lob physical reads 0, lob read-ahead reads 0. SQL Server Execution Times: CPU time = 1420 ms, elapsed time = 1515 ms. The interesting thing for me is that the TASK_REQUEST table is a small table (3 rows at present) and statistics are up to date on the table. Any idea why such different execution plans and execution times would be occuring? And ideally how to change things so that I don't need to use the temp table to get decent performance? The only real difference in the execution plans is that the temp table version introduces an index spool (eager spool) operation.

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• How can I enable a debugging mode via a command-line switch for my Perl program?

  - by Michael Mao

  I am learning Perl in a "head-first" manner. I am absolutely a newbie in this language: I am trying to have a debug_mode switch from CLI which can be used to control how my script works, by switching certain subroutines "on and off". And below is what I've got so far: #!/usr/bin/perl -s -w # purpose : make subroutine execution optional, # which is depending on a CLI switch flag use strict; use warnings; use constant DEBUG_VERBOSE => "v"; use constant DEBUG_SUPPRESS_ERROR_MSGS => "s"; use constant DEBUG_IGNORE_VALIDATION => "i"; use constant DEBUG_SETPPING_COMPUTATION => "c"; our ($debug_mode); mainMethod(); sub mainMethod # () { if(!$debug_mode) { print "debug_mode is OFF\n"; } elsif($debug_mode) { print "debug_mode is ON\n"; } else { print "OMG!\n"; exit -1; } checkArgv(); printErrorMsg("Error_Code_123", "Parsing Error at..."); verbose(); } sub checkArgv #() { print ("Number of ARGV : ".(1 + $#ARGV)."\n"); } sub printErrorMsg # ($error_code, $error_msg, ..) { if(defined($debug_mode) && !($debug_mode =~ DEBUG_SUPPRESS_ERROR_MSGS)) { print "You can only see me if -debug_mode is NOT set". " to DEBUG_SUPPRESS_ERROR_MSGS\n"; die("terminated prematurely...\n") and exit -1; } } sub verbose # () { if(defined($debug_mode) && ($debug_mode =~ DEBUG_VERBOSE)) { print "Blah blah blah...\n"; } } So far as I can tell, at least it works...: the -debug_mode switch doesn't interfere with normal ARGV the following commandlines work: ./optional.pl ./optional.pl -debug_mode ./optional.pl -debug_mode=v ./optional.pl -debug_mode=s However, I am puzzled when multiple debug_modes are "mixed", such as: ./optional.pl -debug_mode=sv ./optional.pl -debug_mode=vs I don't understand why the above lines of code "magically works". I see both of the "DEBUG_VERBOS" and "DEBUG_SUPPRESS_ERROR_MSGS" apply to the script, which is fine in this case. However, if there are some "conflicting" debug modes, I am not sure how to set the "precedence of debug_modes"? Also, I am not certain if my approach is good enough to Perlists and I hope I am getting my feet in the right direction. One biggest problem is that I now put if statements inside most of my subroutines for controlling their behavior under different modes. Is this okay? Is there a more elegant way? I know there must be a debug module from CPAN or elsewhere, but I want a real minimal solution that doesn't depend on any other module than the "default". And I cannot have any control on the environment where this script will be executed...

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• build-helper-maven-plugin add-source does not working when trying to add linked resources

  - by Julian

  I am new to maven and hit a problem that looks easy in the first place but I already kept me busy for a whole day about and no way to get it working. First as part of running eclipse:eclipse plugin I create a linked folder like below: <linkedResources> <linkedResource> <name>properties</name> <type>2</type> <location>${PARENT-2-PROJECT_LOC}/some_other_project/properties</location> </linkedResource> <linkedResource> <name>properties/messages.properties</name> <type>1</type> <location>${PARENT-2-PROJECT_LOC}/some_other_project/properties/messages.properties</location> </linkedResource> And then I am adding that folder as a source folder like below: <plugin> <groupId>org.codehaus.mojo</groupId> <artifactId>build-helper-maven-plugin</artifactId> <version>1.7</version> <executions> <execution> <id>add-source</id> <phase>generate-sources</phase> <goals> <goal>add-source</goal> </goals> <configuration> <sources> <source>properties</source> <source>some_real_folder</source> </sources> </configuration> </execution> </executions> </plugin> However when I am looking at the generated .classpath in eclipse the “some_real_folder” is there but the “properties” is not. It looks like by default the build-helper-maven-plugin will check if the folder is there and if it is not it won’t add it. I am using maven 3.0.4 outside eclipse to run the build and I can see in the maven logs something like this: [INFO] Source directory: <some path>\properties added. This is my project structure: project1 \-- properties (this is the real folder) project2 \-- some_real_folder \-- properties (this is the link resource pointing to the project1/properties folder) All I need is to have both "some_real_folder" and the linked resource "properties" added to the .classpath of the project2

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• SQL Design Question regarding schema and if Name value pair is the best solution

  - by Aur

  I am having a small problem trying to decide on database schema for a current project. I am by no means a DBA. The application parses through a file based on user input and enters that data in the database. The number of fields that can be parsed is between 1 and 42 at the current moment. The current design of the database is entirely flat with there being 42 columns; some have repeated columns such as address1, address2, address3, etc... This says that I should normalize the data. However, data integrity is not needed at this moment and the way the data is shaped I'm looking at several joins. Not a bad thing but the data is still in a 1 to 1 relationship and I still see a lot of empty fields per row. So my concerns are that this does not allow the database or the application to be very extendable. If they want to add more fields to be parsed (which they do) than I'd need to create another table and add another foreign key to the linking table. The third option is I have a table where the fields are defined and a table for each record. So what I was thinking is to make a table that stores the value and then links to those two tables. The problem is I can picture the size of that table growing large depending on the input size. If someone gives me a file with 300,000 records than 300,000 x 40 = 12 million so I have some reservations. However I think if I get to that point than I should be happy it is being used. This option also allows for more custom displaying of information albeit a bit more work but little rework even if you add more fields. So the problem boils down to: 1. Current design is a flat file which makes extending it hard and it is not normalized. 2. Normalize the tables although no real benefits for the moment but requirements change. 3. Normalize it down into the name value pair and hope size doesn't hurt. There are a large number of inserts, updates, and selects against that table. So performance is a worry but I believe the saying is design now, performance testing later? I'm probably just missing something practical so any comments would be appreciated even if it’s a quick sanity check. Thank you for your time.

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• ie8 playing funny with list-style-position: inside

  - by LeeR

  Ok, So problem here... when using list-style-position:inside in IE8 the first like is indented but every line after that is not. So the new lines appear under the bullet. This is fine, but when I use a list with that css applied with an a tag within the li then the text automatically gets pushed to the second line, and the first line is empty. When I remove the a tag from the li then it jumps back up. Any idea on why this might be or is this a bug in the ie8 world or do I just need to double check my css? Any insights would be much appreciated. As asked here is some code <div id="sub_nav"> <ul> ... <li><a class="active_page" href="#">Liposculpture</a> <ul> <li><a href="#">What is Liposculpture?</a></li> <li><a href="#">About Liposculpture surgery</a></li> <li><a href="#" class="active_sub">After Liposculpture surgery</a></li> <li><a href="#">Post Op Instructions</a></li> <li><a href="#">Liposculpture Side Effects</a></li> <li><a href="#">Liposuction Introduction to</a></li> <li><a href="#">Tumescent Liposculpture</a></li> </ul> </li> ... </ul> </div> For the CSS I will try and show it best I can #sub_nav li { width: 200px; padding:4px 0; border-bottom: 1px #CCC solid; } #sub_nav li a { text-decoration: none; color:#555; padding:7px 15px 7px 15px; display: block; } #sub_nav li ul li { list-style-position: inside; list-style-type: disc; font: 11px Arial; padding-left:15px; color:#FFF; border-bottom: none; } #sub_nav li ul li a { padding:0; margin:0; text-indent: 0; } Hope this helps

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• Why am I getting ClassNotFoundExpection when I have properly imported said class and am looking at it in its directory?

  - by Strider

  This is my Javac compiling statement: javac -cp "C:\java\code\j3D\j3dcore.jar;C:\java\code\j3D\j3dutils.jar;C:\java\code\j3D\vecmath.jar" Simple.java compiles with no problems. The three jar files (j3dcore, j3dutils, and vecmath) are the essential jar's for my program (or at least I am led to believe according to this official tutorial on J3D For the record I ripped this code almost line from line from this pdf file. jar files are correctly located in referenced locations When I run my Simple program, (java Simple) I am greeted with Exception in thread "main" java.lang.NoClassDefFoundError: javax/media/j3d/Cavas3d Caused by: java.lang.ClassNotFoundExpection: javax.media.j3d.Canvas3D Currently I am staring directly at this Canvas3D.class that is located within j3dcore.jar\javax\media\j3d\ wtfisthis.jpg Here is the source code: //First java3D Program import java.applet.Applet; import java.awt.BorderLayout; import java.awt.Frame; import java.awt.event.*; import com.sun.j3d.utils.applet.MainFrame; import com.sun.j3d.utils.universe.*; import com.sun.j3d.utils.geometry.ColorCube; import javax.media.j3d.*; import javax.vecmath.*; import java.awt.GraphicsConfiguration; public class Simple extends Applet { public Simple() { setLayout(new BorderLayout()); GraphicsConfiguration config = SimpleUniverse.getPreferredConfiguration(); Canvas3D canvas3D = new Canvas3D(config); add("Center", canvas3D); BranchGroup scene = createSceneGraph(); scene.compile(); // SimpleUniverse is a Convenience Utility class SimpleUniverse simpleU = new SimpleUniverse(canvas3D); // This moves the ViewPlatform back a bit so the // objects in the scene can be viewed. simpleU.getViewingPlatform().setNominalViewingTransform(); simpleU.addBranchGraph(scene); } // end of HelloJava3Da (constructor) public BranchGroup createSceneGraph() { // Create the root of the branch graph BranchGroup objRoot = new BranchGroup(); // Create a simple shape leaf node, add it to the scene graph. // ColorCube is a Convenience Utility class objRoot.addChild(new ColorCube(0.4)); return objRoot; } public static void main(String args[]){ Simple world = new Simple(); } }` Did I import correctly? Did I incorrectly reference my jar files in my Javac statement? If I clearly see Canvas3D within its correct directory why cant java find it? The first folder in both j3dcore.jar and vecmath.jar is "javax". Is the compiler getting confused? If the compiler is getting confused how do I specify where to find that exact class when referencing it within my source code?

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• When to use new layouts and when to use new activities?

  - by cmdfrg

  I'm making a game in Android and I'm trying to add a set of menu screens. Each screen takes up the whole display and has various transitions available to other screens. As a rough summary, the menu screens are: Start screen Difficult select screen Game screen. Pause screen. Game over screen. And there are several different ways you can transition between screen: 1 - 2 2 - 3 3 - 4 (pause game) 4 - 1 (exit game) 4 - 3 (resume game) 3 - 5 (game ends) Obviously, I need some stored state when moving between screens, such as the difficulty level select when starting a game and what the player's score is when the game over screen is shown. Can anyone give me some advice for the easiest way to implement the above screens and transitions in Android? All the create/destroy/pause/resume methods make me nervous about writing brittle code if I'm not careful. I'm not fond of using an Activity for each screen. It seems too heavy weight, having to pass data around using intents seems like a real pain and each screen isn't a useful module by itself. As the "back" button doesn't always go back to the previous screen either, my menu layout doesn't seem to fit the activity model well. At the moment, I'm representing each screen as an XML layout file and I have one activity. I set the different buttons on each layout to call setContentView to update the screen the main activity is showing (e.g. the pause button changes the layout to the pause screen). The activity holds onto all the state needed (e.g. the current difficulty level and the game high score), which makes it easy to share data between screens. This seems roughly similar to the LunarLander sample, except I'm using multiple screens. Does what I have at the moment sound OK or am I not doing things the typical Android way? Is there a class I can use (e.g. something like ViewFlipper) that could make my life easier? By the way, my game screen is implemented as a SurfaceView that stores the game state. I need the state in this view to persist between calls to setContentView (e.g. to resume from paused). Is the right idea to create the game view when the activity starts, keep a reference to it and then use this reference with setContentView whenever I want the game screen to appear?

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• Best way to ask confirmation from user before leaving the page

  - by JohnathanKong

  Hey Everyone, I am currently building a registration page where if the user leaves, I want to pop up a CSS box asking him if he is sure or not. I can accomplish this feat using confirm boxes, but the client says that they are too ugly. I've tried using unload and beforeunload, but both cannot stop the page from being redirected. Using those to events, I return false, so maybe there's a way to cancel other than returning false? Another solution that I've had was redirecting them to another page that has my popup, but the problem with that is that if they do want to leave the page, and it wasn't a mistake, they lose the page they were originally trying to go to. If I was a user, that would irritate me. The last solution was real popup window. The only thing I don't like about that is that the main winow will have their destination page while the pop will have my page. In my opinion it looks disjoint. On top of that, I'd be worried about popup blockers. Just to add to everyones comments. I understand that it is irritating to prevent users from exiting the page, and in my opinion it should not be done. Right now I am using a confirm box at this point. What happens is that it's not actually "preventing" the user from leaving, what the client actually wants to do is make a suggestion if the user is having doubts about registering. If the user is halfway through the registraiton process and leaves for some reason, the client wants to offer the user a free coupon to a seminar (this client is selling seminars) to hopefully persuade the user to register. The client is under the impression that since the user is already on the form, he is thinking of registering, and therefore maybe a seminar of what he is registering for would be the final push to get the user to register. Ideally I don't have to prevent the user from leaving, what would be just as good, and in my opinion better is if I can pause the unload process. Maybe a sleep command? I don't really have to keep the user on the page because either way they will be leaving to go to a different page. Also, as people have stated, this is a terriable title, so if someone knows a better one, I'd really appreciate it if they could change the title to something no so spammer inviting.

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• Why is Dictionary.First() so slow?

  - by Rotsor

  Not a real question because I already found out the answer, but still interesting thing. I always thought that hash table is the fastest associative container if you hash properly. However, the following code is terribly slow. It executes only about 1 million iterations and takes more than 2 minutes of time on a Core 2 CPU. The code does the following: it maintains the collection todo of items it needs to process. At each iteration it takes an item from this collection (doesn't matter which item), deletes it, processes it if it wasn't processed (possibly adding more items to process), and repeats this until there are no items to process. The culprit seems to be the Dictionary.Keys.First() operation. The question is why is it slow? Stopwatch watch = new Stopwatch(); watch.Start(); HashSet<int> processed = new HashSet<int>(); Dictionary<int, int> todo = new Dictionary<int, int>(); todo.Add(1, 1); int iterations = 0; int limit = 500000; while (todo.Count > 0) { iterations++; var key = todo.Keys.First(); var value = todo[key]; todo.Remove(key); if (!processed.Contains(key)) { processed.Add(key); // process item here if (key < limit) { todo[key + 13] = value + 1; todo[key + 7] = value + 1; } // doesn't matter much how } } Console.WriteLine("Iterations: {0}; Time: {1}.", iterations, watch.Elapsed); This results in: Iterations: 923007; Time: 00:02:09.8414388. Simply changing Dictionary to SortedDictionary yields: Iterations: 499976; Time: 00:00:00.4451514. 300 times faster while having only 2 times less iterations. The same happens in java. Used HashMap instead of Dictionary and keySet().iterator().next() instead of Keys.First().

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