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• The 80 column limit, still useful?

  - by Tim Post

  Related: While coding, how many columns do you format for? Is there a valid reason for enforcing a maximum width of 80 characters in a code file, this day and age? I mostly use C, however this question is language agnostic. Its also subjective, so I'll tag it as such. Many individual projects set their own various coding standards, a guide to adjust your coding style. Many enforce an 80 column limit on code, i.e. don't force a dumb 80 x 25 terminal to wrap your lines in someone else's editor of choice if they are stuck with such a display, don't force them to turn off wrapping. Both private and open source projects usually have some style guidelines. My question is, in this day and age, is that requirement more of a pest than a helper? Does anyone still login via the local console with no framebuffer and actually edit code? If so, how often and why cant you use SSH? I help to manage a few open source projects, I was considering extending this limit to 110 columns, but I wanted to get feedback first. So, any feedback is appreciated. I can see the need to make certain OUTPUT of programs (i.e. a --help /h display) 80 columns or less, but I really don't see the need to force people to break up code under 110 columns long into 2 lines, when its easier to read on one line. I can also see the case for adhering to an 80 column limit if you're writing code that will be used on micro controllers that have to be serviced in the field with a god-knows-what terminal emulator. Beyond that, what are your thoughts? Edit: This is not an exact duplicate. I am asking very specific questions, such as how many people are actually still using such a display. I am also not asking "what is a good column limit", I'm proposing one and hoping to gather feedback. Beyond that, I'm also citing cases where the 80 column limit is still a good idea. I don't want a guide to my own "c-style", I'm hoping to adjust standards for several projects. If the duplicate in question had answered all of my questions, I would not have posted this one :) That will teach me to mention it next time. Edit 2 question |= COMMUNITY_WIKI

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• Illegal Instruction When Programming C++ on Linux

  - by remagen

  Heyo, My program, which does exactly the same thing every time it runs (moves a point sprite into the distance) will randomly fail with the text on the terminal 'Illegal Instruction'. My googling has found people encountering this when writing assembly which makes sense because assembly throws those kinds of errors. But why would g++ be generating an illegal instruction like this? It's not like I'm compiling for Windows then running on Linux (which even then, as long as both are on x86 shouldn't AFAIK cause an Illegal Instruction). I'll post the main file below. I can't reliably reproduce the error. Although, if I make random changes (add a space here, change a constant there) that force a recompile I can get a binary which will fail with Illegal Instruction every time it is run, until I try setting a break point, which makes the illegal instruction 'dissapear'. :( #include <stdio.h> #include <stdlib.h> #include <GL/gl.h> #include <GL/glu.h> #include <SDL/SDL.h> #include "Screen.h" //Simple SDL wrapper #include "Textures.h" //Simple OpenGL texture wrapper #include "PointSprites.h" //Simple point sprites wrapper double counter = 0; /* Here goes our drawing code */ int drawGLScene() { /* These are to calculate our fps */ static GLint T0 = 0; static GLint Frames = 0; /* Move Left 1.5 Units And Into The Screen 6.0 */ glLoadIdentity(); glTranslatef(0.0f, 0.0f, -6); glClear(GL_COLOR_BUFFER_BIT | GL_DEPTH_BUFFER_BIT | GL_STENCIL_BUFFER_BIT); glEnable(GL_POINT_SPRITE_ARB); glTexEnvi(GL_POINT_SPRITE, GL_COORD_REPLACE, GL_TRUE); glBegin( GL_POINTS ); /* Drawing Using Triangles */ glVertex3d(0.0,0.0, 0); glVertex3d(1.0,0.0, 0); glVertex3d(1.0,1.0, counter); glVertex3d(0.0,1.0, 0); glEnd( ); /* Finished Drawing The Triangle */ /* Move Right 3 Units */ /* Draw it to the screen */ SDL_GL_SwapBuffers( ); /* Gather our frames per second */ Frames++; { GLint t = SDL_GetTicks(); if (t - T0 >= 50) { GLfloat seconds = (t - T0) / 1000.0; GLfloat fps = Frames / seconds; printf("%d frames in %g seconds = %g FPS\n", Frames, seconds, fps); T0 = t; Frames = 0; counter -= .1; } } return 1; } GLuint objectID; int main( int argc, char **argv ) { Screen screen; screen.init(); screen.resize(800,600); LoadBMP("./dist/Debug/GNU-Linux-x86/particle.bmp"); InitPointSprites(); while(true){drawGLScene();} }

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• Insert Registration Data in MySQL using PHP

  - by J M 4

  I may not be asking this in the best way possible but i will try my hardest. Thank you ahead of time for your help: I am creating an enrollment website which allows an individual OR manager to enroll for medical testing services for professional athletes. I will NOT be using the site as a query DB which anybody can view information stored within the database. The information is instead simply stored, and passed along in a CSV format to our network provider so they can use as needed after the fact. There are two possible scenarios: Scenario 1 - Individual Enrollment If an individual athlete chooses to enroll him/herself, they enter their personal information, submit their payment information (credit/bank account) for processing, and their information is stored in an online database as Athlete1. Scenario 2 - Manager Enrollment If a manager chooses to enroll several athletes he manages/ promotes for, he enters his personal information, then enters the personal information for each athlete he wishes to pay for (name, address, ssn, dob, etc), then submits payment information for ALL athletes he is enrolling. This number can range from 1 single athlete, up to 20 athletes per single enrollment (he can return and complete a follow up enrollment for additional athletes). Initially, I was building the database to house ALL information regardless of enrollment type in a single table which housed over 400 columns (think 20 athletes with over 10 fields per athlete such as name, dob, ssn, etc). Now that I think about it more, I believe create multiple tables (manager(s), athlete(s)) may be a better idea here but still not quite sure how to go about it for the following very important reasons: Issue 1 If I list the manager as the parent table, I am afraid the individual enrolling athlete will not show up in the primary table and will not be included in the overall registration file which needs to be sent on to the network providers. Issue 2 All athletes being enrolled by a manager are being stored in SESSION as F1FirstName, F2FirstName where F1 and F2 relate to the id of the fighter. I am not sure technically speaking how to store multiple pieces of information within the same table under separate rows using PHP. For example, all athleteswill have a first name. The very basic theory of what i am trying to do is: If number_of_athletes 1, store F1FirstName in row 1, column 1 of Table "Athletes"; store F1LastName in row 1, column 2 of Table "Athletes"; store F2FirstName in row 2, column 1 of Table "Athletes"; store F2LastName in row 2, column 2 of table "Athletes"; Does this make sense? I know this question is very long and probably difficult so i appreciate the guidance.

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• MS Access 07 - Q re lookup column vs many-to-many; Q re checkboxes in many-to-many forms

  - by TBinLondon

  Hello, I'm creating a database with Access. This is just a test database, similar to my requirements, so I can get my skills up before creating one for work. I've created a database for a fictional school as this is a good playground and rich data (many students have many subjects have many teachers, etc). Question 1 What is the difference, if any, between using a Lookup column and a many-to-many associate table? Example: I have Tables 'Teacher' and 'Subject'. Many teachers have many subjects. I can, and have, created a table 'Teacher_Subject' and run queries with this. I have then created a lookup column in teachers table with data from subjects. The lookup column seems to take the place of the teacher_subject table. (though the data on relationships is obviously duplicated between lookup table and teacher_subject and may vary). Which one is the 'better' option? Is there a snag with using lookup tables? (I realize that this is a very 'general' question. Links to other resources and answers saying 'that depends...' are appreciated) Question 2 What attracts me to lookup tables is the following: When creating a form for entering subjects for teachers, with lookup I can simply create checkboxes and click a subject for a teacher 'on' or 'off'. Each click on/off creates/removes a record in the lookup column (which replaces teacher_subject). If I use a form from a query from teacher subject with teacher as main form and subject as subform I run into this problem: In the subform I can either select each subject that teacher has in a bombo box, i.e. click, scroll down, select, go to next row, click, scroll down, etc. (takes too long) OR I can create a list box listing all available subjects in each row but allowing me to select only one. (takes up too much space). Is it possible to have a click on/off list box for teacher_subject, creating/removing a record there with each click? Note - I know zero SQL or VB. If the correct answer is "you need to know SQL for this" then that's cool. I just need to know. Thanks!

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• Representation of a DateTime as the local to remote user

  - by TwoSecondsBefore

  Hello! I was confused in the problem of time zones. I am writing a web application that will contain some news with dates of publication, and I want the client to see the date of publication of the news in the form of corresponding local time. However, I do not know in which time zone the client is located. I have three questions. I have to ask just in case: does DateTimeOffset.UtcNow always returns the correct UTC date and time, regardless of whether the server is dependent on daylight savings time? For example, if the first time I get the value of this property for two minutes before daylight savings time (or before the transition from daylight saving time back) and the second time in 2 minutes after the transfer, whether the value of properties in all cases differ by only 4 minutes? Or here require any further logic? (Question #1) Please see the following example and tell me what you think. I posted the news on the site. I assume that DateTimeOffset.UtcNow takes into account the time zone of the server and the daylight savings time, and so I immediately get the correct UTC server time when pressing the button "Submit". I write this value to a MS SQL database in the field of type datetime2(0). Then the user opens a page with news and no matter how long after publication. This may occur even after many years. I did not ask him to enter his time zone. Instead, I get the offset of his current local time from UTC using the javascript function following way: function GetUserTimezoneOffset() { var offset = new Date().getTimezoneOffset(); return offset; } Next I make the calculation of the date and time of publication, which will show the user: public static DateTime Get_Publication_Date_In_User_Local_DateTime( DateTime Publication_Utc_Date_Time_From_DataBase, int User_Time_Zone_Offset_Returned_by_Javascript) { int userTimezoneOffset = User_Time_Zone_Offset_Returned_by_Javascript; // For // example Javascript returns a value equal to -300, which means the // current user's time differs from UTC to 300 minutes. Ie offset // is UTC +6. In this case, it may be the time zone UTC +5 which // currently operates summer time or UTC +6 which currently operates the // standard time. // Right? (Question #2) DateTimeOffset utcPublicationDateTime = new DateTimeOffset(Publication_Utc_Date_Time_From_DataBase, new TimeSpan(0)); // get an instance of type DateTimeOffset for the // date and time of publication for further calculations DateTimeOffset publication_DateTime_In_User_Local_DateTime = utcPublicationDateTime.ToOffset(new TimeSpan(0, - userTimezoneOffset, 0)); return publication_DateTime_In_User_Local_DateTime.DateTime;// return to user } Is the value obtained correct? Is this the right approach to solving this problem? (Question #3)

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• When mocking a class with Moq, how can I CallBase for just specific methods?

  - by Daryn

  I really appreciate Moq's Loose mocking behaviour that returns default values when no expectations are set. It's convenient and saves me code, and it also acts as a safety measure: dependencies won't get unintentionally called during the unit test (as long as they are virtual). However, I'm confused about how to keep these benefits when the method under test happens to be virtual. In this case I do want to call the real code for that one method, while still having the rest of the class loosely mocked. All I have found in my searching is that I could set mock.CallBase = true to ensure that the method gets called. However, that affects the whole class. I don't want to do that because it puts me in a dilemma about all the other properties and methods in the class that hide call dependencies: if CallBase is true then I have to either Setup stubs for all of the properties and methods that hide dependencies -- Even though my test doesn't think it needs to care about those dependencies, or Hope that I don't forget to Setup any stubs (and that no new dependencies get added to the code in the future) -- Risk unit tests hitting a real dependency. Q: With Moq, is there any way to test a virtual method, when I mocked the class to stub just a few dependencies? I.e. Without resorting to CallBase=true and having to stub all of the dependencies? Example code to illustrate (uses MSTest, InternalsVisibleTo DynamicProxyGenAssembly2) In the following example, TestNonVirtualMethod passes, but TestVirtualMethod fails - returns null. public class Foo { public string NonVirtualMethod() { return GetDependencyA(); } public virtual string VirtualMethod() { return GetDependencyA();} internal virtual string GetDependencyA() { return "! Hit REAL Dependency A !"; } // [... Possibly many other dependencies ...] internal virtual string GetDependencyN() { return "! Hit REAL Dependency N !"; } } [TestClass] public class UnitTest1 { [TestMethod] public void TestNonVirtualMethod() { var mockFoo = new Mock<Foo>(); mockFoo.Setup(m => m.GetDependencyA()).Returns(expectedResultString); string result = mockFoo.Object.NonVirtualMethod(); Assert.AreEqual(expectedResultString, result); } [TestMethod] public void TestVirtualMethod() // Fails { var mockFoo = new Mock<Foo>(); mockFoo.Setup(m => m.GetDependencyA()).Returns(expectedResultString); // (I don't want to setup GetDependencyB ... GetDependencyN here) string result = mockFoo.Object.VirtualMethod(); Assert.AreEqual(expectedResultString, result); } string expectedResultString = "Hit mock dependency A - OK"; }

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• Send a variable on the heap to another thread

  - by user1201889

  I have a strange problem in C++. An address of a Boolean gets "destroyed" but it doesn't get touched. I know that there are beater way's to accomplish what I try to do, but I want to know what I do wrong. I have a main class; this main class contains a vector of another class. There is a strange problem when a new instance gets created of this object. This is how my code works: There will start a thread when the constructor gets called of the “2nd”object. This thread gets as Parameter a struct. This is the struct: struct KeyPressData { vector<bool> *AutoPressStatus; vector<int> *AutoPressTime; bool * Destroy; bool * Ready; }; The struct gets filled in the constructor: MultiBoxClient::MultiBoxClient() { //init data DestroyThread = new bool; ReadyThread = new bool; AutoThreadData = new KeyPressData; //Reseting data *DestroyThread = false; *ReadyThread = false; //KeyPressData configurating AutoThreadData->AutoPressStatus = &AutoPressStatus; AutoThreadData->AutoPressTime = &AutoPressTime; AutoThreadData->Destroy = DestroyThread; AutoThreadData->Ready = ReadyThread; //Start the keypress thread CreateThread(NULL,NULL,(LPTHREAD_START_ROUTINE)AutoKeyThread,AutoThreadData,NULL,NULL); } As long as the constructor is running will the program run fine. But when the constructor closes the address of the “AutoThreadData-Destroy” will get corrupted. The program will crash when I call the value of the pointer. void WINAPI AutoKeyThread(void * ThreadData) { KeyPressData * AutoThreadData = (KeyPressData*)ThreadData; while(true) { if(*AutoThreadData->Destroy == true) //CRASH { *AutoThreadData->Ready = true; return; } Sleep(100); } } What did I test: I logged the address of the AutoThreadData and the AutoThreadData-Destroy when the constrcutor is running and clossed; the AutoThreadData address is equal to AutoThreadData when the constructor is closed. So there is no problem here. The address of AutoThreadData-Destroy gets destroyed when the constructor is closed. But how can this happen? The Boolean is on the heap and the KeyPressData struct (AutoThreadData) is on the heap. Destroy before: 00A85328 Destroy after: FEEEFEEE Can someone maby explain why this crash? I know that I can send a pointer to my class to the thread. But I want to know what goes wrong here. That way I can learn from my mistakes. Could someone help me with this problem? Thanks!

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• NHibernateUnitOfWork + ASP.Net MVC

  - by Felipe

  Hi Guys, hows it going? I'm in my first time with DDD, so I'm begginer! So, let's take it's very simple :D I developed an application using asp.net mvc 2 , ddd and nhibernate. I have a domain model in a class library, my repositories in another class library, and an asp.net mvc 2 application. My Repository base class, I have a construct that I inject and dependency (my unique ISessionFactory object started in global.asax), the code is: public class Repository<T> : IRepository<T> where T : Entidade { protected ISessionFactory SessionFactory { get; private set; } protected ISession Session { get { return SessionFactory.GetCurrentSession(); } } protected Repository(ISessionFactory sessionFactory) { SessionFactory = sessionFactory; } public void Save(T entity) { Session.SaveOrUpdate(entity); } public void Delete(T entity) { Session.Delete(entity); } public T Get(long key) { return Session.Get<T>(key); } public IList<T> FindAll() { return Session.CreateCriteria(typeof(T)).SetCacheable(true).List<T>(); } } And After I have the spefic repositories, like this: public class DocumentRepository : Repository<Domain.Document>, IDocumentRepository { // constructor public DocumentRepository (ISessionFactory sessionFactory) : base(sessionFactory) { } public IList<Domain.Document> GetByType(int idType) { var result = Session.CreateQuery("from Document d where d.Type.Id = :IdType") .SetParameter("IdType", idType) .List<Domain.Document>(); return result; } } there is not control of transaction in this code, and it's working fine, but, I would like to make something to control this repositories in my controller of asp.net mvc, something simple, like this: using (var tx = /* what can I put here ? */) { try { _repositoryA.Save(objA); _repositoryB.Save(objB); _repositotyC.Delete(objC); /* ... others tasks ... */ tx.Commit(); } catch { tx.RollBack(); } } I've heared about NHibernateUnitOfWork, but i don't know :(, How Can I configure NHibernateUnitOfWork to work with my repositories ? Should I change the my simple repository ? Sugestions are welcome! So, thanks if somebody read to here! If can help me, I appretiate! PS: Sorry for my english! bye =D

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• AnimationDrawable, when does it end?

  - by Syb

  I know there have been several people with the same question. Which is: How do i know when a frame by frame animation has ended? I have not had any useful answer on fora i visited. So i thought, let's see if they know at stackoverflow. But I could not sit still in the mean time, so i made a work around of this, but it does not really work the way i would like it to. here is the code: public class Main extends Activity { AnimationDrawable sybAnimation; /** Called when the activity is first created. */ @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.main); ImageView imageView = (ImageView)findViewById(R.id.ImageView01); imageView.setBackgroundResource(R.anim.testanimation); sybAnimation = (AnimationDrawable) imageView.getBackground(); imageView.post(new Starter()); } class Starter implements Runnable { public void run() { sybAnimation.start(); long totalDuration = 0; for(int i = 0; i< sybAnimation.getNumberOfFrames();i++){ totalDuration += sybAnimation.getDuration(i); } Timer timer = new Timer(); timer.schedule(new AnimationFollowUpTimerTask(R.id.ImageView01, R.anim.testanimation_reverse),totalDuration); } } class AnimationFollowUpTimerTask extends TimerTask { private int id; private int animationToRunId; public AnimationFollowUpTimerTask(int idOfImageView, int animationXML){ id = idOfImageView; animationToRunId = animationXML; } @Override public void run() { ImageView imageView = (ImageView)findViewById(id); imageView.setBackgroundResource(animationToRunId); AnimationDrawable anim = (AnimationDrawable) imageView.getBackground(); anim.start(); } } basically I make a timertask which is scheduled with the same time as the animation to take. In that run() I want to load a new animation into the imageView and start that animation, this however does not work. Does anyone know how to get this to work, or even better, have a better way to find out when an AnimationDrawable has ended its animation?

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• CoreGraphics taking a while to show on a large view - can i get it to repeat pixels?

  - by Andrew

  This is my coregraphics code: void drawTopPaperBackground(CGContextRef context, CGRect rect) { CGRect paper3 = CGRectMake(10, 14, 300, rect.size.height - 14); CGRect paper2 = CGRectMake(13, 12, 294, rect.size.height - 12); CGRect paper1 = CGRectMake(16, 10, 288, rect.size.height - 10); //Shadow CGContextSetShadowWithColor(context, CGSizeMake(0,0), 10, [[UIColor colorWithWhite:0 alpha:0.5]CGColor]); CGPathRef path = createRoundedRectForRect(paper3, 0); CGContextSetFillColorWithColor(context, [[UIColor blackColor] CGColor]); CGContextAddPath(context, path); CGContextFillPath(context); //Layers of paper //CGContextSaveGState(context); drawPaper(context, paper3); drawPaper(context, paper2); drawPaper(context, paper1); //CGContextRestoreGState(context); } void drawPaper(CGContextRef context, CGRect rect) { //Shadow CGContextSaveGState(context); CGContextSetShadowWithColor(context, CGSizeMake(0,0), 1, [[UIColor colorWithWhite:0 alpha:0.5]CGColor]); CGPathRef path = createRoundedRectForRect(rect, 0); CGContextSetFillColorWithColor(context, [[UIColor blackColor] CGColor]); CGContextAddPath(context, path); CGContextFillPath(context); //CGContextRestoreGState(context); //Gradient //CGContextSaveGState(context); CGColorRef startColor = [UIColor colorWithWhite:0.92 alpha:1.0].CGColor; CGColorRef endColor = [UIColor colorWithWhite:0.94 alpha:1.0].CGColor; CGRect firstHalf = CGRectMake(rect.origin.x, rect.origin.y, rect.size.width / 2, rect.size.height); CGRect secondHalf = CGRectMake(rect.origin.x + (rect.size.width / 2), rect.origin.y, rect.size.width / 2, rect.size.height); drawVerticalGradient(context, firstHalf, startColor, endColor); drawVerticalGradient(context, secondHalf, endColor, startColor); //CGContextRestoreGState(context); //CGContextSaveGState(context); CGRect redRect = rectForRectWithInset(rect, -1); CGMutablePathRef redPath = createRoundedRectForRect(redRect, 0); //CGContextSaveGState(context); CGContextSetStrokeColorWithColor(context, [[UIColor blackColor] CGColor]); CGContextAddPath(context, path); CGContextClip(context); CGContextAddPath(context, redPath); CGContextSetShadowWithColor(context, CGSizeMake(0, 0), 15.0, [[UIColor colorWithWhite:0 alpha:0.1] CGColor]); CGContextStrokePath(context); CGContextRestoreGState(context); } The view is a UIScrollView, which contains a textview. Every time the user types something and goes onto a new line, I call [self setNeedsDisplay]; and it redraws the code. But when the view starts to get long - around 1000 height, it has very noticeable lag. How can i make this code more efficient? Can i take a line of pixels and make it just repeat that, or stretch it, all the way down?

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• BlackBerry OS 7.1 secured TLS connection is closed after very short time

  - by MrVincenzo

  To make a long story short: Same client-server configuration, same network topology, same device (Bold 9900) - works perfectly well on OS 7.0 but doesn't work as expected on OS 7.1 and the secured tls connection is being closed by the server after a very short time. My application opens a secured tls connection to a server. The connection is kept alive by a application layer keep-alive mechanism and remains open until the client closes it. Attached is a simplified version of the actual code that opens connection and reads from the socket. The code works perfectly on OS 5.0-7.0 but doesn't work as expected on OS 7.1. When running on OS 7.1, the blocking read() returns with -1 (end of the stream has been reached) after very short time (10-45 seconds). For OS 5.0-7.0 the call to read() remains blocking until next data arrives and the connection is never closed by the server. Connection connection = Connector.open(connectionString); connInputStream = connection.openInputStream(); while (true) { try { retVal = connInputStream.read(); if (-1 == retVal) { break; // end of stream has been reached } } catch (Exception e ) { // do error handling } // data read from stream is handled here } UPDATE 1: Apparently, the problem appears only when I use secured tls connection (either using mobile network or WiFi) on OS 7.1. Everything works as expected when opening a non secured connection on OS 7.1. For tls on mobile networks I use the following connection string: connectionString = "tls://someipaddress:443;deviceside=false;ConnectionType=mds-public;EndToEndDesired"; For tls on Wifi I use the following connection string: connectionString = "tls://someipaddress:443;deviceside=true;interface=wifi;EndToEndRequired" UPDATE 2: The connection is never idle. I am constantly receiving and sending data on it. The issue appears both when using mobile connection and WiFi. The issue appears both on real OS 7.1 devices and simulators. I am starting to suspect that it is somehow related either to the connection string I am using or to the tls handshake. UPDATE 3: According to Wireshark's captures that I made with the OS 7.1 simulator, the secured tls connection is being closed by the server (client receives FIN). For the moment I don't have the server's private key therefore I unable to debug the tls handshake.

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• SQL Server 2005: Update rows in a specified order (like ORDER BY)?

  - by JMTyler

  I want to update rows of a table in a specific order, like one would expect if including an ORDER BY clause, but SQL Server does not support the ORDER BY clause in UPDATE queries. I have checked out this question which supplied a nice solution, but my query is a bit more complicated than the one specified there. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) ORDER BY Parent.Depth DESC; So, what I'm hoping that you'll notice is that a single table (TableA) contains a hierarchy of rows, wherein one row can be the parent or child of any other row. The rows need to be updated in order from the deepest child up to the root parent. This is because TableA.ColA must contain an up-to-date concatenation of its own current value with the values of its children (I realize this query only concats with one child, but that is for the sake of simplicity - the purpose of the example in this question does not necessitate any more verbosity), therefore the query must update from the bottom up. The solution suggested in the question I noted above is as follows: UPDATE messages SET status=10 WHERE ID in (SELECT TOP (10) Id FROM Table WHERE status=0 ORDER BY priority DESC ); The reason that I don't think I can use this solution is because I am referencing column values from the parent table inside my subquery (see WHERE Child.ParentColB = Parent.ColB), and I don't think two sibling subqueries would have access to each others' data. So far I have only determined one way to merge that suggested solution with my current problem, and I don't think it works. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) WHERE Parent.Id IN (SELECT Id FROM TableA ORDER BY Parent.Depth DESC); The WHERE..IN subquery will not actually return a subset of the rows, it will just return the full list of IDs in the order that I want. However (I don't know for sure - please tell me if I'm wrong) I think that the WHERE..IN clause will not care about the order of IDs within the parentheses - it will just check the ID of the row it currently wants to update to see if it's in that list (which, they all are) in whatever order it is already trying to update... Which would just be a total waste of cycles, because it wouldn't change anything. So, in conclusion, I have looked around and can't seem to figure out a way to update in a specified order (and included the reason I need to update in that order, because I am sure I would otherwise get the ever-so-useful "why?" answers) and I am now hitting up Stack Overflow to see if any of you gurus out there who know more about SQL than I do (which isn't saying much) know of an efficient way to do this. It's particularly important that I only use a single query to complete this action. A long question, but I wanted to cover my bases and give you guys as much info to feed off of as possible. :) Any thoughts?

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• I have to do two seemingly mutually exclusive things on leaving an asp:textbox. Please help me get

  - by aape

  This project has gone from being a simple '99 Ford F-150 to the Homer. I've got controls with a gridview with textboxes for data entry. All the user controls on the pages are in AJAX updatepanels. User types in a database column or budget entity or some other financial thing they want to include in the report. The textboxes in the gridview have autopostback = true set. overly long background info When the user leaves the textbox, during the postback (triggered by onTextChanged) I do some validation back on the server on their entry - regexs, do they have rights to that column, is that column locked, etc. If it fails, I put a error message next to the textbox. If it passes, I wipe out any title or error that used to be next to the code. Focus is getting lost from the postback if they're tabbing out of the box, rather than going to the next textbox in the gridview. So to fix that I need, if their leaving the tb via the tab key, to also figure out what textbox or gridviewrow they're on, if they're not on the last row, and after the validation and labeling, put the focus on the textbox in the next row. I can't figure out how, in ontextchanged, to find what caused me to leave the textbox, so I'm thinking use javascript onkeyup to test the key pressed and then find the next box etc, but the ontextchanged fires first and then the js never does, and also, since the control is all AJAXed, the javascript can't find the textboxes because when you enter the page everything is collapsed (the requirements people loooove to collapse and expand things), and so when it's expanded, all the 'new' textboxes are up in the viewstate stuff in the page source, and not down where javascript can see them. The questions So I'm wondering if I can have an onblur in the javascript that can trigger a postback where I can do my validation and such, and either 1) include the keypressed or pick it out of sender in the event or 2) followup the onblur with onkeyup and somehow figure out what textbox is next on the grid and throw focus there. Or, is there another .NET based approach that could work for this? In terms of tearing the whole thing down and starting from scratch, I couldn't sell that to the bosses, I'm past the point of no return as far as that goes.

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• Are there compelling reasons not to use Groovy?

  - by Leonard H Martin

  I'm developing a LoB application in Java after a long absence from the platform (having spent the last 8 years or so entrenched in Fortran, C, a smidgin of C++ and latterly .Net). Java, the language, is not much changed from how I remember it. I like it's strengths and I can work around its weaknesses - the platform has grown and deciding upon the myriad of different frameworks which appear to do much the same thing as one another is a different story; but that can wait for another day - all-in-all I'm comfortable with Java. However, over the last couple of weeks I've become enamoured with Groovy, and purely from a selfish point of view: but not just because it makes development against the JVM a more succinct and entertaining (and, well, "groovy") proposition than Java (the language). What strikes me most about Groovy is its inherent maintainability. We all (I hope!) strive to write well documented, easy to understand code. However, sometimes the languages we use themselves defeat us. An example: in 2001 I wrote a library in C to translate EDIFACT EDI messages into ANSI X12 messages. This is not a particularly complicated process, if slightly involved, and I thought at the time I had documented the code properly - and I probably had - but some six years later when I revisited the project (and after becoming acclimatised to C#) I found myself lost in so much C boilerplate (mallocs, pointers, etc. etc.) that it took three days of thoughtful analysis before I finally understood what I'd been doing six years previously. This evening I've written about 2000 lines of Java (it is the day of rest, after all!). I've documented as best as I know how, but, but, of those 2000 lines of Java a significant proportion is Java boiler plate. This is where I see Groovy and other dynamic languages winning through - maintainability and later comprehension. Groovy lets you concentrate on your intent without getting bogged down on the platform specific implementation; it's almost, but not quite, self documenting. I see this as being a huge boon to me when I revisit my current project (which I'll port to Groovy asap) in several years time and to my successors who will inherit it and carry on the good work. So, are there any reasons not to use Groovy?

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• What is the proper way to wrap this php loop around the javascript

  - by zac

  Hi, I am mashing together the Google Maps 2 script with a Wordpress loop so there is a CMS platform for the map data. I have this working fine : var point = new GLatLng(48.5139,-123.150531); var marker = createMarker(point,"Lime Kiln State Park", '<?php $post_id = 182; $my_post = get_post($post_id); $mapTitle = $my_post->post_title; $mapIMG = get_post_meta($post_id, 'mapImage', true); $snip = get_post_meta($post_id, 'mapExcerpt', true); echo "<div class=\"span-12\">"; echo "<div class=\"mapTitle\">"; echo $mapTitle; echo "</div>"; echo "<img class=\"mapImage\" src=\""; echo bloginfo('url'); echo "/wp-content/files_mf/"; echo $mapIMG; echo "\" /> "; echo "<div class=\"mapContent\">"; echo $snip; echo "</div>"; echo "<div class=\"moreLink\">"; echo "<a href=\""; echo $permalink = get_permalink( $post_id ); echo "\">Find out more &raquo; </a>"; echo "</div>"; echo "</div>"; ?>') map.addOverlay(marker); However I am hoping to also be able to include the two variables at the start within the php loop so that both of those can also be generated by custom fields. Can someone please show me that the proper way to write this would be so that all of the data can be pulled in from fields within that post id? So the lat/long and title could also be set from within the post 182 fields.

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• WPF PathGeometry/RotateTransform optimization

  - by devinb

  I am having performance issues when rendering/rotating WPF triangles If I had a WPF triangle being displayed and it will be rotated to some degree around a centrepoint, I can do it one of two ways: Programatically determine the points and their offset in the backend, use XAML to simply place them on the canvas where they belong, it would look like this: <Path Stroke="Black"> <Path.Data> <PathGeometry> <PathFigure StartPoint ="{Binding CalculatedPointA, Mode=OneWay}"> <LineSegment Point="{Binding CalculatedPointB, Mode=OneWay}" /> <LineSegment Point="{Binding CalculatedPointC, Mode=OneWay}" /> <LineSegment Point="{Binding CalculatedPointA, Mode=OneWay}" /> </PathFigure> </PathGeometry> </Path.Data> </Path> Generate the 'same' triangle every time, and then use a RenderTransform (Rotate) to put it where it belongs. In this case, the rotation calculations are being obfuscated, because I don't have any access to how they are being done. <Path Stroke="Black"> <Path.Data> <PathGeometry> <PathFigure StartPoint ="{Binding TriPointA, Mode=OneWay}"> <LineSegment Point="{Binding TriPointB, Mode=OneWay}" /> <LineSegment Point="{Binding TriPointC, Mode=OneWay}" /> <LineSegment Point="{Binding TriPointA, Mode=OneWay}" /> </PathFigure> </PathGeometry> </Path.Data> <Path.RenderTransform> <RotateTransform CenterX="{Binding Centre.X, Mode=OneWay}" CenterY="{Binding Centre.Y, Mode=OneWay}" Angle="{Binding Orientation, Mode=OneWay}" /> </Path.RenderTransform> </Path> My question is which one is faster? I know I should test it myself but how do I measure the render time of objects with such granularity. I would need to be able to time how long the actual rendering time is for the form, but since I'm not the one that's kicking off the redraw, I don't know how to capture the start time.

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• ODBC: when is the best time to create my database?

  - by mawg

  I have a windows program which generates PGP forms which will be filled in later. Those PHP forms will populate a database. It looks very much like MySql, but I can't be certain, so let's call it ODBC. And, yes, it does have to be a windows program. There will also be PHP forms which query the database - examine which tables and fields it contains and then generates forms which can be used to search the database (e.g, it finds a table with fields "employee_name", etc and generates a form which lets you search based on employee name. Let's call that design time and run time. At design time, some manager or IT guy or similar gets to define the nature of the database and at runtime 1) a worker fills in the form daily and 2) management can extract reports. Here's my question: given that the database is defined at "design time" (and populated at run time), where and how is best to do so? 1 I could use an ODBC interface from the windows program, but I am having difficulty finding something good to work with Delphi. Things like ADO and firebird tend to expect you to already have a database and allow you to manipulate it, but I can find no code example of how to create a database and some tables, so ... 2 I could used DOS commands from Delphi in my windows program. I just tried and got a response to MySql --version, but am not sure if MySql etc are more interactive. That is, can I use a script file or a very long stacked command with semicolons and returns separating? e.g 'CREATE DATABASE db; CREATE TABLE t1;' 3) Since the best way to work with databases seems to be PHP, perhaps my windows program could spit out a PHP page which would, when run in a browser, create the database. I have tried to make this as uncomplicated as I can, but please feel free to ask questions. It may be that there are several valid ways, but there is probably one 'better' solution in terms of ease of implementation or maintenance. Better scratch option 3. What if the user later wants to come back and have the windows program change the input form? It needs to update the database too.

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• Best Practices for Working with Multiple Monitors in Visual Studio 2010

  - by Clever Human

  Now that Visual Studio 2010 has support for multiple monitors, I am curious how other people have their environments arranged. I have yet to come up with an arrangement that I am really satisfied with. The current best I have come up with for my 2 monitor system is to have all code windows detached. Then, on my primary monitor, I am able to have two code windows side by side (using the Windows 7 keyboard shortcuts WinKey+LeftArrow and WinKey+RightArrow.) On my secondary monitor I put the rest of the IDE with all of the tool windows that are normally on the bottom (errors list, find window, call stack, etc...) docked where the code windows normally go. I've also tried having all those things detached and having almost nothing in the IDE proper. The problems with this layout are: Newly opened code windows always open in the IDE, not on top of one of the detached windows. Detached code windows do not remember their exact placement from session to session (they are slightly off, having me to use the winkey + arrow key shortcut again and again for each window. There seems to be no way to have the code panes aware that they are on top of one another (IE -- multiple tabs.) The CTRL+TAB shortcut always displays on top of the IDE proper. The Code Panes are always "on top" of (children of) the IDE. So clicking on any code pane brings the IDE to the foreground, even when I care only about that code pane, and not the IDE. Other more minor issues... What would go a long way to making this better is having the code panes detach such that they are tab strips that can have other code panes docked within them. The new multi-monitor support in VS2010 is good, but it still seems really lacking. Can these issues be solved with an add-in? If so, is anyone aware of one? Is there a better way to work with the IDE on multiple monitors than what I am doing? NOTE: While this question is subjective (there is certainly no "this is the best way and that's final" answer) I'd really like to know possibly better methods of working with the IDE than what I have come up with. The intent is not to start a "mine's best" flame war.

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• Are there any modern GUI toolkits which implement a heirarchical menu buffer zone?

  - by scomar

  In Bruce Tognazzini's quiz on Fitt's Law, the question discussing the bottleneck in the hierarchical menu (as used in almost every modern desktop UI), talks about his design for the original Mac: The bottleneck is the passage between the first-level menu and the second-level menu. Users first slide the mouse pointer down to the category menu item. Then, they must carefully slide the mouse directly across (horizontally) in order to move the pointer into the secondary menu. The engineer who originally designed hierarchicals apparently had his forearm mounted on a track so that he could move it perfectly in a horizontal direction without any vertical component. Most of us, however, have our forarms mounted on a pivot we like to call our elbow. That means that moving our hand describes an arc, rather than a straight line. Demanding that pivoted people move a mouse pointer along in a straight line horizontally is just wrong. We are naturally going to slip downward even as we try to slide sideways. When we are not allowed to slip downward, the menu we're after is going to slam shut just before we get there. The Windows folks tried to overcome the pivot problem with a hack: If they see the user move down into range of the next item on the primary menu, they don't instantly close the second-level menu. Instead, they leave it open for around a half second, so, if users are really quick, they can be inaccurate but still get into the second-level menu before it slams shut. Unfortunately, people's reactions to heightened chance of error is to slow down, rather than speed up, a well-established phenomenon. Therefore, few users will ever figure out that moving faster could solve their problem. Microsoft's solution is exactly wrong. When I specified the Mac hierarchical menu algorthm in the mid-'80s, I called for a buffer zone shaped like a <, so that users could make an increasingly-greater error as they neared the hierarchical without fear of jumping to an unwanted menu. As long as the user's pointer was moving a few pixels over for every one down, on average, the menu stayed open, no matter how slow they moved. (Cancelling was still really easy; just deliberately move up or down.) This just blew me away! Such a simple idea which would result in a huge improvement in usability. I'm sure I'm not the only one who regularly has the next level of a menu slam shut because I don't move the mouse pointer in a perfectly horizontal line. So my question is: Are there any modern UI toolkits which implement this brilliant idea of a < shaped buffer zone in hierarchical menus? And if not, why not?!

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• Please help with very simple android widget button click. Getting very frustrated. :(

  - by Little Wendy

  I have poured over every example that I can find and I have gone through the official documentation from google. All I am trying to do is create a toast notification from a button click on a widget. I can get the widget (and button) to appear but I can't get my intent to fire. I have seen several examples that show doing this different ways but I just can't get it to work. I haven't felt this helpless with coding in a long time and I'm starting to feel dumb. This is what I have so far: public class simpleclick extends AppWidgetProvider { /** Called when the activity is first created. */ @Override public void onUpdate(Context context, AppWidgetManager appWidgetManager, int[] appWidgetIds) { final int N = appWidgetIds.length; Toast.makeText(context, "doing on update", Toast.LENGTH_SHORT).show(); for (int i=0; i<N; i++) { int appWidgetId = appWidgetIds[i]; Intent intent = new Intent(context, simpleclick.class); intent.setAction("ham"); PendingIntent pendingIntent = PendingIntent.getActivity(context, 0, intent, 0); RemoteViews views = new RemoteViews(context.getPackageName(), R.layout.main); views.setOnClickPendingIntent(R.id.Timm, pendingIntent); appWidgetManager.updateAppWidget(appWidgetId, views); } } //@Override public void onReceive(Context context, Intent intent) { // TODO Auto-generated method stub Toast.makeText(context, "action is: " + intent.getAction(), Toast.LENGTH_SHORT).show(); super.onReceive(context, intent); } } My manifest: <?xml version="1.0" encoding="utf-8"?> <manifest xmlns:android="http://schemas.android.com/apk/res/android" package="com.tblabs.simpleclick" android:versionCode="1" android:versionName="1.0"> <application android:icon="@drawable/icon" android:label="@string/app_name"> <receiver android:name="simpleclick"> <intent-filter> <action android:name="android.appwidget.action.APPWIDGET_UPDATE" /> </intent-filter> <meta-data android:name="android.appwidget.provider" android:resource="@xml/simpleclick" /> </receiver> </application> <uses-sdk android:minSdkVersion="5" /> </manifest> I would appreciate some help! Thanks, Wendy

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• java.lang.NumberFormatException: unable to parse '' as integer one more time

  - by Quzziy

  I will take two numbers from user, but this number from EditText must be converted to int. I think it should be working, but I still have problem with compilation code in Android Studio. CatLog show error in line with: int wiek = Integer.parseInt(wiekEditText.getText().toString()); Below is my full Android code: public class MyActivity extends ActionBarActivity { int Wynik; @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_my); int Tmax, RT; EditText wiekEditText = (EditText) findViewById(R.id.inWiek); EditText tspoczEditText = (EditText) findViewById(R.id.inTspocz); int wiek = Integer.parseInt(wiekEditText.getText().toString()); int tspocz = Integer.parseInt(tspoczEditText.getText().toString()); Tmax = 220 - wiek; RT = Tmax - tspocz; Wynik = 70*RT/100 + tspocz; final EditText tempWiekEdit = wiekEditText; TabHost tabHost = (TabHost) findViewById(R.id.tabHost); //Do TabHost'a z layoutu tabHost.setup(); TabHost.TabSpec tabSpec = tabHost.newTabSpec("Calc"); tabSpec.setContent(R.id.Calc); tabSpec.setIndicator("Calc"); tabHost.addTab(tabSpec); tabSpec = tabHost.newTabSpec("Hints"); tabSpec.setContent(R.id.Hints); tabSpec.setIndicator("Hints"); tabHost.addTab(tabSpec); final Button Btn = (Button) findViewById(R.id.Btn); Btn.setOnClickListener(new View.OnClickListener() { @Override public void onClick(View v) { Toast.makeText(getApplicationContext(),"blablabla"+ "Wynik",Toast.LENGTH_SHORT).show(); } }); wiekEditText.addTextChangedListener(new TextWatcher() { @Override public void beforeTextChanged(CharSequence s, int start, int count, int after) { } @Override public void onTextChanged(CharSequence s, int start, int before, int count) { Btn.setEnabled(!(tempWiekEdit.getText().toString().trim().isEmpty())); } @Override public void afterTextChanged(Editable s) { } }); } @Override public boolean onCreateOptionsMenu(Menu menu) { // Inflate the menu; this adds items to the action bar if it is present. getMenuInflater().inflate(R.menu.my, menu); return true; } @Override public boolean onOptionsItemSelected(MenuItem item) { // Handle action bar item clicks here. The action bar will // automatically handle clicks on the Home/Up button, so long // as you specify a parent activity in AndroidManifest.xml. int id = item.getItemId(); if (id == R.id.action_settings) { return true; } return super.onOptionsItemSelected(item); } }

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• PHP Based session variable not retaining value. Works on localhost, but not on server.

  - by Foo

  I've been trying to debug this problem for many hours, but to no avail. I've been using PHP for many years, and got back into it after long hiatus, so I'm still a bit rusty. Anyways, my $_SESSION vars are not retaining their value for some reason that I can't figure out. The site worked on localhost perfectly, but uploading it to the server seemed to break it. First thing I checked was the PHP.ini server settings. Everything seems fine. In fact, my login system is session based and it works perfectly. So now that I know $_SESSIONS are working properly and retaining their value for my login, I'm presuming the server is setup and the problem is in my script. Here's a stripped version of the code that's causing a problem. $type, $order and $style are not being retained after they are set via a GET variable. The user clicks a link, which sets a variable via GET, and this variable is retained for the remainder of their session. Is there some problem with my logic that I'm not seeing? <?php require_once('includes/top.php'); //first line includes a call to session_start(); require_once('includes/db.php'); $type = filter_input(INPUT_GET, 't', FILTER_VALIDATE_INT); $order = filter_input(INPUT_GET, 'o', FILTER_VALIDATE_INT); $style = filter_input(INPUT_GET, 's', FILTER_VALIDATE_INT); /* According to documentation, filter_input returns a NULL when variables are undefined. So, if t, o, or s are not set via URL, $type, $order and $style will be NULL. */ print_r($_SESSION); /* All other sessions, such as the login session, etc. are displayed here. After the sessions are set below, they are displayed up here to... simply with no value. This leads me to believe the problem is with the code below, perhaps? */ // If $type is not null (meaning it WAS set via the get method above) // or it's false because the validation failed for some reason, // then set the session to the $type. I removed the false check for simplicity. // This code is being successfully executed, and the data is being stored... if(!is_null($type)) { $_SESSION['type'] = $type; } if(!is_null($order)) { $_SESSION['order'] = $order; } if(!is_null($style)) { $_SESSION['style'] = $style; } $smarty->display($template); ?> If anyone can point me in the right direction, I'd greatly appreciate it. Thanks.

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• sifr menu list problems, height of object calculated wrong

  - by armyofda12mnkeys

  I am using Drupal and have sifr set on list items, and also a, a:hover are set via Drupal so that the links hover. I looked at the sifr3-rules.js file drupal's render module creates and it looks good. and in fact my other sifr items look fine... But the list item ones are goofing for some reason... There is extra space below the list, so if i have a list item, and inside of that, I have a unordered list with more list items, the Flash Object made in the 1st li (which will cover the rest of the sublist items too), is too bid so you see space under the children until the next parent li comes up. (so looks like extra bottom padding on that portion of the list in IE8... in FF almost similar but each subitem has space at bottom... with javascript turned off, you see the list items look fine by themselves).... Also if the parent list item is shorter than the sub list items text, the width for the flash object is set only as long as the first list item, and therefore cuts off the rest the sublist item's text. Any idea how to resolve any of these? Only unusual thing i see i am doing is setting forceSingleLine and preventWrap (which dont make a difference if taken off). ****Edit, I may just try to figure out how to get Drupal's menu-block module to output a around my hyperlinks in the list items... then i can target the div's with my rule (and the a,a:hover rules will apply), so each menu item gets its own sifr object instead of sifr3 trying to figure out how to do the lists and sublists. Very useful to me is anyone knows a way to target a hyperlink (<a> tag) and also allow a:hover rules. I know how to do it with pretend another tag that includes hyperlinks, like if i had <h2><a>sometitle</a></h2>, i could have a rule for h2, but then use the a, a:hover rules in the sifr3-rules.js file to target that. so i would need a way to target the hyperlink in the list, but also apply a :hover to it (not sure if this can be done since its not underneith for example the h2 tag).

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• Not seeing Sync Block in Object Layout

  - by bob-bedell

  It's my understanding the all .NET object instances begin with an 8 byte 'object header': a synch block (4 byte pointer into a SynchTableEntry table), and a type handle (4 byte pointer into the types method table). I'm not seeing this in VS 2010 RC's (CLR 4.0) debugger memory windows. Here's a simple class that will generate a 16 byte instance, less the object header. class Program { short myInt = 2; // 4 bytes long myLong = 3; // 8 bytes string myString = "aString"; // 4 byte object reference // 16 byte instance static void Main(string[] args) { new Program(); return; } } An SOS object dump tells me that the total object size is 24 bytes. That makes sense. My 16 byte instance plus an 8 byte object header. !DumpObj 0205b660 Name: Offset_Test.Program MethodTable: 000d383c EEClass: 000d13f8 Size: 24(0x18) bytes File: C:\Users\Bob\Desktop\Offset_Test\Offset_Test\bin\Debug\Offset_Test.exe Fields: MT Field Offset Type VT Attr Value Name 632020fc 4000001 10 System.Int16 1 instance 2 myInt 632050d8 4000002 4 System.Int64 1 instance 3 myLong 631fd2b8 4000003 c System.String 0 instance 0205b678 myString Here's the raw memory: 0x0205B660 000d383c 00000003 00000000 0205b678 00000002 ... And here are some annotations: offset 0 000d383c ;TypeHandle (pointer to MethodTable), 4 bytes offset 4 00000003 00000000 ;myLong, 8 bytes offset 12 0205b678 ;myString, 4 byte reference to address of "myString" on GC Heap offset 16 00000002 ;myInt, 4 bytes My object begins a address 0x0205B660. But I can only account for 20 bytes of it, the type handle and the instance fields. There is no sign of a synch block pointer. The object size is reported as 24 bytes, but the debugger is showing that it only occupies 20 bytes of memory. I'm reading Drill Into .NET Framework Internals to See How the CLR Creates Runtime Objects, and expected the first 4 bytes of my object to be a zeroed synch block pointer, as shown in Figure 8 of that article. Granted, this is an article about CLR 1.1. I'm just wondering if the difference between what I'm seeing and what this early article reports is a change in either the debugger's display of object layout, or in the way the CLR lays out objects in versions later than 1.1. Anyway, can anyone account for my 4 missing bytes?

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• How to use an adjacency matrix to determine which rows to 'pass' to a function in r?

  - by dubhousing

  New to R, and I have a long-ish question: I have a shapefile/map, and I'm aiming to calculate a certain index for every polygon in that map, based on attributes of that polygon and each polygon that neighbors it. I have an adjacency matrix -- which I think is the same as a "1st-order queen contiguity weights matrix", although I'm not sure -- that describes which polygons border which other polygons, e.g., POLYID A B C D E A 0 0 1 0 1 B 0 0 1 0 0 C 1 1 0 1 0 D 0 0 1 0 1 E 1 0 0 1 0 The above indicates, for instance, that polygons 'C' and 'E' adjoin polygon 'A'; polygon 'B' adjoins only polygon 'C', etc. The attribute table I have has one polygon per row: POLYID TOT L10K 10_15K 15_20K ... A 500 24 30 77 ... Where TOT, L10K, etc. are the variables I use to calculate an index. There are 525 polygons/rows in my data, so I'd like to use the adjacency matrix to determine which rows' attributes to incorporate into the calculation of the index of interest. For now, I can calculate the index when I subset the rows that correspond to one 'bundle' of neighboring polygons, and then use a loop (if it's of interest, I'm calculating the Centile Gap Index, a measure of local income segregation). E.g., subsetting the 'neighborhood' of the Detroit City Schools: Detroit <- UNSD00[c(142,150,164,221,226,236,295,327,157,177,178,364,233,373,418,424,449,451,487),] Then record the marginal column proportions and a running total: catprops <- vector() for(i in 4:19) { catprops[(i-3)]<-sum(Detroit[,i])/sum(Detroit[,3]) } catprops <- as.data.frame(catprops) catprops[,2]<-cumsum(catprops[,1]) Columns 4:19 are the necessary ones in the attribute table. Then I use the following code to calculate the index -- note that the loop has "i in 1:19" because the Detroit subset has 19 polygons. cgidistsum <- 0 for(i in 1:19) { pranks <- vector() for(j in 4:19) { if (Detroit[i,j]==0) pranks <- append(pranks,0) else if (j == 4) pranks <- append(pranks,seq(0,catprops[1,2],by=catprops[1,2]/Detroit[i,j])) else pranks <- append(pranks,seq(catprops[j-4,2],catprops[j-3,2],by=catprops[j-3,1]/Detroit[i,j])) } distpranks <- vector() distpranks<-abs(pranks-median(pranks)) cgidistsum <- cgidistsum + sum(distpranks) } cgi <- (.25-(cgidistsum/sum(Detroit[,3])))/.25 My apologies if I've provided more information than is necessary. I would really like to exploit the adjacency matrix in order to calculate the CGI for each 'bundle' of these rows. If you happen to know how I could started with this, that would be great. and my apologies for any novice mistakes, I'm new to R!

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