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  • Problem on jboss lookup entitymanager

    - by Stefano
    I have my ear-project deployed in jboss 5.1GA. From webapp i don't have problem, the lookup of my ejb3 work fine! es: ShoppingCart sc= (ShoppingCart) (new InitialContext()).lookup("idelivery-ear-1.0/ShoppingCartBean/remote"); also the iniection of my EntityManager work fine! @PersistenceContext private EntityManager manager; From test enviroment (I use Eclipse) the lookup of the same ejb3 work fine! but the lookup of entitymanager or PersistenceContext don't work!!! my good test case: public void testClient() { Properties properties = new Properties(); properties.put("java.naming.factory.initial","org.jnp.interfaces.NamingContextFactory"); properties.put("java.naming.factory.url.pkgs","org.jboss.naming:org.jnp.interfaces"); properties.put("java.naming.provider.url","localhost"); Context context; try{ context = new InitialContext(properties); ShoppingCart cart = (ShoppingCart) context.lookup("idelivery-ear-1.0/ShoppingCartBean/remote"); // WORK FINE } catch (Exception e) { e.printStackTrace(); } } my bad test : EntityManagerFactory emf = Persistence.createEntityManagerFactory("idelivery"); EntityManager em = emf.createEntityManager(); //test1 EntityManager em6 = (EntityManager) new InitialContext().lookup("java:comp/env/persistence/idelivery"); //test2 PersistenceContext em3 = (PersistenceContext)(new InitialContext()).lookup("idelivery/remote"); //test3 my persistence.xml <persistence-unit name="idelivery" transaction-type="JTA"> <jta-data-source>java:ideliveryDS</jta-data-source> <properties> <property name="hibernate.hbm2ddl.auto" value="create-drop" /><!--validate | update | create | create-drop--> <property name="hibernate.dialect" value="org.hibernate.dialect.MySQL5InnoDBDialect" /> <property name="hibernate.show_sql" value="true" /> <property name="hibernate.format_sql" value="true" /> </properties> </persistence-unit> my datasource: <datasources> <local-tx-datasource> <jndi-name>ideliveryDS</jndi-name> ... </local-tx-datasource> </datasources> I need EntityManager and PersistenceContext to test my query before build ejb... Where is my mistake?

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  • How can I load a file into a DataBag from within a Yahoo PigLatin UDF?

    - by Cervo
    I have a Pig program where I am trying to compute the minimum center between two bags. In order for it to work, I found I need to COGROUP the bags into a single dataset. The entire operation takes a long time. I want to either open one of the bags from disk within the UDF, or to be able to pass another relation into the UDF without needing to COGROUP...... Code: # **** Load files for iteration **** register myudfs.jar; wordcounts = LOAD 'input/wordcounts.txt' USING PigStorage('\t') AS (PatentNumber:chararray, word:chararray, frequency:double); centerassignments = load 'input/centerassignments/part-*' USING PigStorage('\t') AS (PatentNumber: chararray, oldCenter: chararray, newCenter: chararray); kcenters = LOAD 'input/kcenters/part-*' USING PigStorage('\t') AS (CenterID:chararray, word:chararray, frequency:double); kcentersa1 = CROSS centerassignments, kcenters; kcentersa = FOREACH kcentersa1 GENERATE centerassignments::PatentNumber as PatentNumber, kcenters::CenterID as CenterID, kcenters::word as word, kcenters::frequency as frequency; #***** Assign to nearest k-mean ******* assignpre1 = COGROUP wordcounts by PatentNumber, kcentersa by PatentNumber; assignwork2 = FOREACH assignpre1 GENERATE group as PatentNumber, myudfs.kmeans(wordcounts, kcentersa) as CenterID; basically my issue is that for each patent I need to pass the sub relations (wordcounts, kcenters). In order to do this, I do a cross and then a COGROUP by PatentNumber in order to get the set PatentNumber, {wordcounts}, {kcenters}. If I could figure a way to pass a relation or open up the centers from within the UDF, then I could just GROUP wordcounts by PatentNumber and run myudfs.kmeans(wordcount) which is hopefully much faster without the CROSS/COGROUP. This is an expensive operation. Currently this takes about 20 minutes and appears to tack the CPU/RAM. I was thinking it might be more efficient without the CROSS. I'm not sure it will be faster, so I'd like to experiment. Anyway it looks like calling the Loading functions from within Pig needs a PigContext object which I don't get from an evalfunc. And to use the hadoop file system, I need some initial objects as well, which I don't see how to get. So my question is how can I open a file from the hadoop file system from within a PIG UDF? I also run the UDF via main for debugging. So I need to load from the normal filesystem when in debug mode. Another better idea would be if there was a way to pass a relation into a UDF without needing to CROSS/COGROUP. This would be ideal, particularly if the relation resides in memory.. ie being able to do myudfs.kmeans(wordcounts, kcenters) without needing the CROSS/COGROUP with kcenters... But the basic idea is to trade IO for RAM/CPU cycles. Anyway any help will be much appreciated, the PIG UDFs aren't super well documented beyond the most simple ones, even in the UDF manual.

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  • DCVS + hosting for a startup commercial multiplatform phone app

    - by AG
    I'm in lean startup mode, working on a simple phone app that will be published initially as a iThingy app and an Android app with, possibly, Blackberry and Symbian versions to follow. I'm about to go from no repository to needing a central repository that up to 4 very part-time resources will be sharing. Two of us have no version control background, one has used Subversion, and I've used most of the major centralized VCS systems. I'm not going to be pushing the technical limitations of any VCS for a long time; I'm sure that any of the major systems would work fine. And the hosting accounts I've looked at seem reasonable. So I'm really focussed on minimizing the downside risks. That is, I'd like to find a stable setup that is easy to learn in general, easy to use from Windows/Eclipse, and won't paint me into any obvious corners for the next 12 months or so. A quick search of the web had led me to consider the following pairs of DVCS and hosting service, with what I think I'm hearing as their strengths and weaknesses (for my purposes): Bazaar/Launchpad -- My initial choice since I need to get more familiar with this pair for the Google Summer of Code mentoring I'm doing. But, whatever the technical merits, a non-starter for me because they are purely open source, no private repositories plans to purchase that I can see. Git/GitHub -- Git: Fast, light, ultimately flexible, but relatively less Windows friendly, Eclipse plugin (eGit) available but relatively young, GitHub: widely used, pricing is fine Mercurial/BitBucket -- Mercurial: a little less flexible, a little more Windows friendly, Eclipse plugin seems a bit more mature, BitBucket: widely used, pricing is fine, includes a wiki and an issue tracker that we might be able to use instead of something like BaseCamp, at least for a while. Mercurial/BitBucket seem like the winning pair so far for my particular situation; at least two of us are definitely going to be working mostly from Eclipse on Windows and reducing my own learning curve is a priority. ;-) But I have two specific questions: 1) Am I wrong about Bazaar/Launchpad and is there a viable, secure way to use them for proprietary code? 2) Any reason to think that the Mecurial/Bitbucket pair will end up being a headache for my Mac developer, soon, or for Blackberry or Symbian developers a little later? ag

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  • How to generate and encode (for use in GA), random, strict, binary rooted trees with N leaves?

    - by Peter Simon
    First, I am an engineer, not a computer scientist, so I apologize in advance for any misuse of nomenclature and general ignorance of CS background. Here is the motivational background for my question: I am contemplating writing a genetic algorithm optimizer to aid in designing a power divider network (also called a beam forming network, or BFN for short). The BFN is intended to distribute power to each of N radiating elements in an array of antennas. The fraction of the total input power to be delivered to each radiating element has been specified. Topologically speaking, a BFN is a strictly binary, rooted tree. Each of the (N-1) interior nodes of the tree represents the input port of an unequal, binary power splitter. The N leaves of the tree are the power divider outputs. Given a particular power divider topology, one is still free to map the power divider outputs to the array inputs in an arbitrary order. There are N! such permutations of the outputs. There are several considerations in choosing the desired ordering: 1) The power ratio for each binary coupler should be within a specified range of values. 2) The ordering should be chosen to simplify the mechanical routing of the transmission lines connecting the power divider. The number of ouputs N of the BFN may range from, say, 6 to 22. I have already written a genetic algorithm optimizer that, given a particular BFN topology and desired array input power distribution, will search through the N! permutations of the BFN outputs to generate a design with compliant power ratios and good mechanical routing. I would now like to generalize my program to automatically generate and search through the space of possible BFN topologies. As I understand it, for N outputs (leaves of the binary tree), there are $C_{N-1}$ different topologies that can be constructed, where $C_N$ is the Catalan number. I would like to know how to encode an arbitrary tree having N leaves in a way that is consistent with a chromosomal description for use in a genetic algorithm. Also associated with this is the need to generate random instances for filling the initial population, and to implement crossover and mutations operators for this type of chromosome. Any suggestions will be welcome. Please minimize the amount of CS lingo in your reply, since I am not likely to be acquainted with it. Thanks in advance, Peter

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  • How can I configure a Factory with the possible providers?

    - by Jonathas Costa
    I have three assemblies: "Framework.DataAccess", "Framework.DataAccess.NHibernateProvider" and "Company.DataAccess". Inside the assembly "Framework.DataAccess", I have my factory (with the wrong implementation of discovery): public class DaoFactory { private static readonly object locker = new object(); private static IWindsorContainer _daoContainer; protected static IWindsorContainer DaoContainer { get { if (_daoContainer == null) { lock (locker) { if (_daoContainer != null) return _daoContainer; _daoContainer = new WindsorContainer(new XmlInterpreter()); // THIS IS WRONG! THIS ASSEMBLY CANNOT KNOW ABOUT SPECIALIZATIONS! _daoContainer.Register( AllTypes.FromAssemblyNamed("Company.DataAccess") .BasedOn(typeof(IReadDao<>)).WithService.FromInterface(), AllTypes.FromAssemblyNamed("Framework.DataAccess.NHibernateProvider") .BasedOn(typeof(IReadDao<>)).WithService.Base()); } } return _daoContainer; } } public static T Create<T>() where T : IDao { return DaoContainer.Resolve<T>(); } } This assembly also defines the base interface for data access IReadDao: public interface IReadDao<T> { IEnumerable<T> GetAll(); } I want to keep this assembly generic and with no references. This is my base data access assembly. Then I have the NHibernate provider's assembly, which implements the above IReadDao using NHibernate's approach. This assembly references the "Framework.DataAccess" assembly. public class NHibernateDao<T> : IReadDao<T> { public NHibernateDao() { } public virtual IEnumerable<T> GetAll() { throw new NotImplementedException(); } } At last, I have the "Company.DataAccess" assembly, which can override the default implementation of NHibernate provider and references both previously seen assemblies. public interface IProductDao : IReadDao<Product> { Product GetByName(string name); } public class ProductDao : NHibernateDao<Product>, IProductDao { public override IEnumerable<Product> GetAll() { throw new NotImplementedException("new one!"); } public Product GetByName(string name) { throw new NotImplementedException(); } } I want to be able to write... IRead<Product> dao = DaoFactory.Create<IRead<Product>>(); ... and then get the ProductDao implementation. But I can't hold inside my base data access any reference to specific assemblies! My initial idea was to read that from a xml config file. So, my question is: How can I externally configure this factory to use a specific provider as my default implementation and my client implementation?

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  • generated service mock: everything but RhinoMocks fails?

    - by hko
    I have the "quest" to search for the next Mocking Framework for my company, and basically it's down to NSubstitute (simplest syntax, but no strict mocks), FakeItEasy(best reviews, Roy Osherove bonus, and slightly better lib support than NSubstitute), Moq (best "other libs support", biggest featureset, downside: mock.Object). We definitely want to move on from RhinoMocks, e.g. because of the unusefull interactiontest error messages (it should tell me what the parameter was instead, when a verification fails). So I was pretty surprised the other day (that was yesterday) when I found out RhinoMocks could do a thing where every other mock framework fails at: Mocking an autogenerated SomethingService (a typical VS autogenerated service with a default construtor in a partial class). Please don't argue about the design.. I intend to write lightweight integration tests (and some unit tests), and I can't mess around with the service, the product is installed on too many customers system. See this code: // here the NSubstitute and FakeItEasy equivalents throw an exception.. see below TicketStoreService fakeTicketStoreService = MockRepository.GenerateMock<TicketStoreService>(); fakeTicketStoreService.Expect(service => service.DoSomething(Arg.Is(new Guid())).Return(new Guid()); fakeTicketStoreService.DoSomething(Arg.Is(new Guid())); fakeTicketStoreService.VerifyAllExpectations(); Note that DoSomething is a non-virtual methodcall in an autogenerated class. So it shouldn't work, according to common knowledge. But it does. Problem is that it's the only (non commercial) framework that can do this: Rhino.Mocks works, and verification works too FakeItEasy says it doesn't find a default constructor (probably just wrong exception message): No default constructor was found on the type SomeNamespace.TicketStoreService Moq gives something sane and understandable: Invalid setup on a non-virtual (overridable in VB) member: service=> service.DoSomething Nsubstitute gives a message System.NotSupportedException: Cannot serialize member System.ComponentModel.Component.Site of type System.ComponentModel.ISite because it is an interface. I'm really wondering what's going on here with the frameworks, except Moq. The "fancy new" frameworks seem to have an initial perf hit too, probably preparing some Type cache and serializing stuff, whilst RhinoMocks somehow manages to create a very "slim" mock without recursion. I have to admit I didn't like RhinoMocks very well, but here it shines.. unfortunately. So, is there a way to get that to work with newer (non-commercial!) mocking frameworks, or somehow get a sane error message out of Rhino.Mocks? And why can Rhino.Mocks achieve this, when clearly every Mocking framework states it can only work with virtual methods when given a concrete class? Let's not derail the discussion by talking about alternative approaches like Extract&Override or runtime-proxy Mocking frameworks like JustMock/TypeMock/Moles or the new Fakes framework, I know these, but that would be less ideal solutions, for reasons beyond this topic. Any help appreciated..

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  • Constructor or Explicit cast

    - by Felan
    In working with Linq to Sql I create a seperate class to ferry data to a web page. To simplify creating these ferry objects I either use a specialized constructor or an explicit conversion operator. I have two questions. First which approach is better from a readibility perspective? Second while the clr code that is generated appeared to be the same to me, are there situations where one would be treated different than the other by the compiler (in lambda's or such). Example code (DatabaseFoo uses specialized constructor and BusinessFoo uses explicit operator): public class DatabaseFoo { private static int idCounter; // just to help with generating data public int Id { get; set; } public string Name { get; set; } public DatabaseFoo() { Id = idCounter++; Name = string.Format("Test{0}", Id); } public DatabaseFoo(BusinessFoo foo) { this.Id = foo.Id; this.Name = foo.Name; } } public class BusinessFoo { public int Id { get; set; } public string Name { get; set; } public static explicit operator BusinessFoo(DatabaseFoo foo) { return FromDatabaseFoo(foo); } public static BusinessFoo FromDatabaseFoo(DatabaseFoo foo) { return new BusinessFoo {Id = foo.Id, Name = foo.Name}; } } public class Program { static void Main(string[] args) { Console.WriteLine("Creating the initial list of DatabaseFoo"); IEnumerable<DatabaseFoo> dafoos = new List<DatabaseFoo>() { new DatabaseFoo(), new DatabaseFoo(), new DatabaseFoo(), new DatabaseFoo(), new DatabaseFoo(), new DatabaseFoo()}; foreach(DatabaseFoo dafoo in dafoos) Console.WriteLine(string.Format("{0}\t{1}", dafoo.Id, dafoo.Name)); Console.WriteLine("Casting the list of DatabaseFoo to a list of BusinessFoo"); IEnumerable<BusinessFoo> bufoos = from x in dafoos select (BusinessFoo) x; foreach (BusinessFoo bufoo in bufoos) Console.WriteLine(string.Format("{0}\t{1}", bufoo.Id, bufoo.Name)); Console.WriteLine("Creating a new list of DatabaseFoo by calling the constructor taking BusinessFoo"); IEnumerable<DatabaseFoo> fufoos = from x in bufoos select new DatabaseFoo(x); foreach(DatabaseFoo fufoo in fufoos) Console.WriteLine(string.Format("{0}\t{1}", fufoo.Id, fufoo.Name)); } }

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  • WPF: Order of stretch sizing

    - by RFBoilers
    I'm creating a modal dialog window which contains three essential parts: a TextBlock containing instructions, a ContentControl for the dialog panel, and a ContentControl for the dialog buttons. Each of these parts are contained in a separate Grid row. I have some specific constraints when it comes to how the dialog should be sized. The issue I'm having is with the instructions TextBlock. I want the instructions to be as wide as the ContentControl for the dialog panel. The instructions should then wrap and grow vertically as needed. Should the instructions not be able to grow vertically, then it should begin to grow horizontally. Getting the instructions to be the width of the ContentControl and grow vertically was simple. The part I can't seem to figure out is how to get it to grow horizontally when out of vertical space. My initial thought was to create a class that extends TextBlock and override MeasureOverride. However, that method is sealed. Currently, I'm playing with the idea of have the dialog Window override MeasureOverride to calculate the available size for the instructions block. Am I missing a much simpler way of accomplishing this? Does anyone have any better ideas than this? Messing with MeasureOverride seems like it will be a lot of work. Here is some sample code to give you a general idea of how the dialog is laid out: <Window x:Class="Dialogs.DialogWindow" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" x:Name="dialogWindow" ShowInTaskbar="False" WindowStyle="None" AllowsTransparency="True" Background="Transparent" ResizeMode="NoResize" SizeToContent="WidthAndHeight" WindowStartupLocation="CenterScreen"> <Border Style="{StaticResource WindowBorderStyle}" Margin="15"> <Grid> <Grid.RowDefinitions> <RowDefinition Height="Auto"/> <RowDefinition Height="Auto"/> <RowDefinition Height="Auto"/> </Grid.RowDefinitions> <TextBlock Margin="25,5" VerticalAlignment="Top" HorizontalAlignment="Left" Text="{Binding Instructions}" TextWrapping="Wrap" Width="{Binding ElementName=panelContentControl, Path=ActualWidth, Mode=OneWay}"/> <ContentControl x:Name="panelContentControl" Grid.Row="1" Margin="25,5" Content="{Binding PanelContent}"/> <ContentControl x:Name="buttonsContentControl" Grid.Row="2" HorizontalAlignment="Right" VerticalAlignment="Center" Margin="25,5" Content="{Binding ButtonsContent}"/> </Grid> </Border> </Window>

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  • c++ quick sort running time

    - by chnet
    I have a question about quick sort algorithm. I implement quick sort algorithm and play it. The elements in initial unsorted array are random numbers chosen from certain range. I find the range of random number effects the running time. For example, the running time for 1, 000, 000 random number chosen from the range (1 - 2000) takes 40 seconds. While it takes 9 seconds if the 1,000,000 number chosen from the range (1 - 10,000). But I do not know how to explain it. In class, we talk about the pivot value can effect the depth of recursion tree. For my implementation, the last value of the array is chosen as pivot value. I do not use randomized scheme to select pivot value. int partition( vector<int> &vec, int p, int r) { int x = vec[r]; int i = (p-1); int j = p; while(1) { if (vec[j] <= x){ i = (i+1); int temp = vec[j]; vec[j] = vec[i]; vec[i] = temp; } j=j+1; if (j==r) break; } int temp = vec[i+1]; vec[i+1] = vec[r]; vec[r] = temp; return i+1; } void quicksort ( vector<int> &vec, int p, int r) { if (p<r){ int q = partition(vec, p, r); quicksort(vec, p, q-1); quicksort(vec, q+1, r); } } void random_generator(int num, int * array) { srand((unsigned)time(0)); int random_integer; for(int index=0; index< num; index++){ random_integer = (rand()%10000)+1; *(array+index) = random_integer; } } int main() { int array_size = 1000000; int input_array[array_size]; random_generator(array_size, input_array); vector<int> vec(input_array, input_array+array_size); clock_t t1, t2; t1 = clock(); quicksort(vec, 0, (array_size - 1)); // call quick sort int length = vec.size(); t2 = clock(); float diff = ((float)t2 - (float)t1); cout << diff << endl; cout << diff/CLOCKS_PER_SEC <<endl; }

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  • Unselecting RadioButtons in Java Swing

    - by Thomas
    When displaying a group of JRadioButtons, initially none of them is selected (unless you programmatically enforce that). I would like to be able to put buttons back into that state even after the user already selected one, i.e., none of the buttons should be selected. However, using the usual suspects doesn't deliver the required effect: calling 'setSelected(false)' on each button doesn't work. Interestingly, it does work when the buttons are not put into a ButtonGroup - unfortunately, the latter is required for JRadioButtons to be mutually exclusive. Also, using the setSelected(ButtonModel, boolean) - method of javax.swing.ButtonGroup doesn't do what I want. I've put together a small program to demonstrate the effect: two radio buttons and a JButton. Clicking the JButton should unselect the radio buttons so that the window looks exactly as it does when it first pops up. import java.awt.Container; import java.awt.GridLayout; import java.awt.event.*; import javax.swing.*; /** * This class creates two radio buttons and a JButton. Initially, none * of the radio buttons is selected. Clicking on the JButton should * always return the radio buttons into that initial state, i.e., * should disable both radio buttons. */ public class RadioTest implements ActionListener { /* create two radio buttons and a group */ private JRadioButton button1 = new JRadioButton("button1"); private JRadioButton button2 = new JRadioButton("button2"); private ButtonGroup group = new ButtonGroup(); /* clicking this button should unselect both button1 and button2 */ private JButton unselectRadio = new JButton("Unselect radio buttons."); /* In the constructor, set up the group and event listening */ public RadioTest() { /* put the radio buttons in a group so they become mutually * exclusive -- without this, unselecting actually works! */ group.add(button1); group.add(button2); /* listen to clicks on 'unselectRadio' button */ unselectRadio.addActionListener(this); } /* called when 'unselectRadio' is clicked */ public void actionPerformed(ActionEvent e) { /* variant1: disable both buttons directly. * ...doesn't work */ button1.setSelected(false); button2.setSelected(false); /* variant2: disable the selection via the button group. * ...doesn't work either */ group.setSelected(group.getSelection(), false); } /* Test: create a JFrame which displays the two radio buttons and * the unselect-button */ public static void main(String[] args) { JFrame frame = new JFrame(); frame.setDefaultCloseOperation(JFrame.EXIT_ON_CLOSE); RadioTest test = new RadioTest(); Container contentPane = frame.getContentPane(); contentPane.setLayout(new GridLayout(3,1)); contentPane.add(test.button1); contentPane.add(test.button2); contentPane.add(test.unselectRadio); frame.setSize(400, 400); frame.setVisible(true); } } Any ideas anyone? Thanks!

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  • javascript simple object creation test: opera leaks?

    - by joe
    Hi, I am trying to figure out certain memory leak conditions in javascript on a few browsers. Currently I'm only testing FF 3.6, Opera 10.10, and Safari 4.0.3. I've started with a fairly simple test, and can confirm no memory leaks in Firefox and Safari. But Opera just takes memory and never gives it back. What gives? Here's the test: <html> <head> <script type="text/javascript"> window.onload = init; //window.onunload = cleanup; var a=[]; function init() { var d = document.createElement('div'); d.innerHTML = "page loading..."; document.body.appendChild(d); for (var i=0; i<400000; i++) { a[i] = new Obj("xxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxxx"); } d.innerHTML = "PAGE LOADED"; } function cleanup() { for (var i=0; i<400000; i++) { a[i] = null; } } function Obj(msg) { this.msg=msg; } </script> </head> <body> </body> </html> I shouldn't need the cleanup() call on window.unload, but tried that also. No luck. As you can see this is simple JS, no circular DOM links, no closures. I monitor the memory usage using 'top' on Mac 10.4.11. Memory usage spikes up on page load, as expected. In FF and Safari reloading the page does not use any further memory, and all memory is returned when the window (tab) is closed. In Opera, memory spikes on load, and seems to also spike further on each reload (but not always...). But regardless of reload, memory never goes back down below the initial load spike. I had hoped this was a no-brainer test that all browsers would pass, so I could move on to more "interesting" conditions. Am I doing something wrong here? Or is this a known Opera issue? Thanks! -joe

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  • Pruning data for better viewing on loglog graph - Matlab

    - by Geodesic
    Hi Guys, just wondering if anyone has any ideas about an issue I'm having. I have a fair amount of data that needs to be displayed on one graph. Two theoretical lines that are bold and solid are displayed on top, then 10 experimental data sets that converge to these lines are graphed, each using a different identifier (eg the + or o or a square etc). These graphs are on a log scale that goes up to 1e6. The first few decades of the graph (< 1e3) look fine, but as all the datasets converge ( 1e3) it's really difficult to see what data is what. There's over 1000 data points points per decade which I can prune linearly to an extent, but if I do this too much the lower end of the graph will suffer in resolution. What I'd like to do is prune logarithmically, strongest at the high end, working back to 0. My question is: how can I get a logarithmically scaled index vector rather than a linear one? My initial assumption was that as my data is lenear I could just use a linear index to prune, which lead to something like this (but for all decades): //%grab indicies per decade ind12 = find(y >= 1e1 & y <= 1e2); indlow = find(y < 1e2); indhigh = find(y > 1e4); ind23 = find(y >+ 1e2 & y <= 1e3); ind34 = find(y >+ 1e3 & y <= 1e4); //%We want ind12 indexes in this decade, find spacing tot23 = round(length(ind23)/length(ind12)); tot34 = round(length(ind34)/length(ind12)); //%grab ones to keep ind23keep = ind23(1):tot23:ind23(end); ind34keep = ind34(1):tot34:ind34(end); indnew = [indlow' ind23keep ind34keep indhigh']; loglog(x(indnew), y(indnew)); But this causes the prune to behave in a jumpy fashion obviously. Each decade has the number of points that I'd like, but as it's a linear distribution, the points tend to be clumped at the high end of the decade on the log scale. Any ideas on how I can do this?

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  • Zend Table Relationship Modeling with Composite Key

    - by emeraldjava
    I have a table with a composite primary key using four columns. mysql> describe leaguesummary; +------------------------+------------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +------------------------+------------------+------+-----+---------+----------------+ | leagueid | int(10) unsigned | NO | PRI | NULL | auto_increment | | leaguetype | enum('I','T') | NO | PRI | NULL | | | leagueparticipantid | int(10) unsigned | NO | PRI | NULL | | | leaguestandard | int(10) unsigned | NO | | NULL | | | leaguedivision | varchar(5) | NO | PRI | NULL | | | leagueposition | int(10) unsigned | NO | | NULL | | I have the league object modelled as so (all plain enough mappings) <?php class Model_DbTable_League extends Zend_Db_Table_Abstract { protected $_name = 'league'; protected $_primary = 'id'; protected $_dependentTables = array('Model_DbTable_LeagueSummary'); And I've started like this on the new model class. I've mapped a simple reference map which returns all rows linked to the league id. // http://files.zend.com/help/Zend-Framework/zend.db.table.relationships.html // http://naneau.nl/2007/04/21/a-zend-framework-tutorial-part-one/ class Model_DbTable_LeagueSummary extends Zend_Db_Table_Abstract { protected $_name = "leaguesummary"; protected $_primary = array('leagueid', 'leaguetype','leagueparticipantid','leaguedivision'); protected $_referenceMap = array( 'Summary' => array( 'columns' => array('leagueid'), 'refTableClass' => 'Model_DbTable_League', 'refColumns' => array('id') ), ..... ); } ?> The simple case works when called from my controller public function listAction() { // action body $leagueTable = new Model_DbTable_League(); $this->view->leagues = $leagueTable->getLeagues(); $league = $leagueTable->getLeague(6); // work $summary = $league->findDependentRowset('Model_DbTable_LeagueSummary','Summary'); Zend_Debug::dump($summary,"",true); I'm not sure how i can define extra _referenceMap keys which will take extra contraint ket values. I would like to be able to define a set called 'MenA' in which the type and division values are hardcoded, and the league id is taken from the initial rowset. 'MenA' =>array( 'columns' => array('leagueid','leaguetype','leaguedivision'), 'refTableClass' => 'Model_DbTable_League', 'refColumns' => array("id","I","A") ) Is this style of mapping possible ie hardcoding the values into the 'refColumns'. The second crazy idea i had was to pass the variable values in as part of the third param of the findDependentRowset() method. $menA = $league->findDependentRowset('Model_DbTable_LeagueSummary','MenA',array("I","A")); Any suggestions on how I might use the Zend DB Table Relationship mapping correctly to do this would be appreciated. I'm not interested in the plain, old and ugly $db-select(a,b,c)-where(..) style solution.

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  • Help with css selector for jquery

    - by NachoF
    I have this wordpress blog that has many pages with subpages that dropdown on hover.... the thing is, I dont want the pages that you hover to link to anything unless they dont have any ul with many anchors inside so just the subpages will have a href different than "#" So basically Im hacking my way through this with some simple javascript. jQuery(document).ready(function(){ jQuery("#menus li > a").attr("href","#"); }); This is selecting every a.. and I dont want that... I just want the anchors that are main pages, not subpages... heres the html so maybe you can think of a better way to select this. Ill explain first the structure is an ul with many li that have an anchor inside if the li also has a ul inside then those are subpages that will appear on hover. hence the initial anchor should have href="#" if there is no ul inside the li then the li a should keep its href. <ul id="menus"> <li> <a href="somelink">Main Page</a> //href should be changed to # <ul> <li> <a href="somelink2/">Subpage1</a> </li> <li> <a href="somelink3">Subpage2</a> </li> </ul> </li> <li> <a href="somelink">MainPage-with-no-subpages</a> //href should not be changed </li> <li> <a href="somelink4">MainPage</a> //href should be changed to # <ul> <li> <a href="somelink5">Subpage</a> </li> <li> <a href="somelink6">Subpage</a> </li> </ul> </li> </ul>

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  • How do I refactor these two C# functions to abstract their logic from the specific class properties

    - by ObligatoryMoniker
    I have two functions whose underlying logic is the same but in one case it sets one property value on a class and in another case it sets a different one. How can I rewrite the following two functions to abstract away as much of the algorithm as possible so that I can make changes in logic in a single place? SetBillingAddress private void SetBillingAddress(OrderAddress newBillingAddress) { BasketHelper basketHelper = new BasketHelper(SiteConstants.BasketName); OrderAddress oldBillingAddress = basketHelper.Basket.Addresses[basketHelper.BillingAddressID]; bool NewBillingAddressIsNotOldBillingAddress = ((oldBillingAddress == null) || (newBillingAddress.OrderAddressId != oldBillingAddress.OrderAddressId)); bool BillingAddressHasBeenPreviouslySet = (oldBillingAddress != null); bool BillingAddressIsNotSameAsShippingAddress = (basketHelper.ShippingAddressID != basketHelper.BillingAddressID); bool NewBillingAddressIsNotShippingAddress = (newBillingAddress.OrderAddressId != basketHelper.ShippingAddressID); if (NewBillingAddressIsNotOldBillingAddress && BillingAddressHasBeenPreviouslySet && BillingAddressIsNotSameAsShippingAddress) { basketHelper.Basket.Addresses.Remove(oldBillingAddress); } if (NewBillingAddressIsNotOldBillingAddress && NewBillingAddressIsNotShippingAddress) { basketHelper.Basket.Addresses.Add(newBillingAddress); } basketHelper.BillingAddressID = newBillingAddress.OrderAddressId; basketHelper.Basket.Save(); } And here is the second one: SetShippingAddress private void SetBillingAddress(OrderAddress newShippingAddress) { BasketHelper basketHelper = new BasketHelper(SiteConstants.BasketName); OrderAddress oldShippingAddress = basketHelper.Basket.Addresses[basketHelper.ShippingAddressID]; bool NewShippingAddressIsNotOldShippingAddress = ((oldShippingAddress == null) || (newShippingAddress.OrderAddressId != oldShippingAddress.OrderAddressId)); bool ShippingAddressHasBeenPreviouslySet = (oldShippingAddress != null); bool ShippingAddressIsNotSameAsBillingAddress = (basketHelper.ShippingAddressID != basketHelper.BillingAddressID); bool NewShippingAddressIsNotBillingAddress = (newShippingAddress.OrderAddressId != basketHelper.BillingAddressID); if (NewShippingAddressIsNotOldShippingAddress && ShippingAddressHasBeenPreviouslySet && ShippingAddressIsNotSameAsBillingAddress) { basketHelper.Basket.Addresses.Remove(oldShippingAddress); } if (NewShippingAddressIsNotOldShippingAddress && NewShippingAddressIsNotBillingAddress) { basketHelper.Basket.Addresses.Add(newShippingAddress); } basketHelper.ShippingAddressID = newShippingAddress.OrderAddressId; basketHelper.Basket.Save(); } My initial thought was that if I could pass a class's property by refernce then I could rewrite the previous functions into something like private void SetPurchaseOrderAddress(OrderAddress newAddress, ref String CurrentChangingAddressIDProperty) and then call this function and pass in either basketHelper.BillingAddressID or basketHelper.ShippingAddressID as CurrentChangingAddressIDProperty but since I can't pass C# properties by reference I am not sure what to do with this code to be able to reuse the logic in both places. Thanks for any insight you can give me.

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  • .NET unit test runner outputting FaultException.Detail

    - by Adam
    Hello, I am running some unit tests on a WCF service. The service is configured to include exception details in the fault response (with the following in my service configuration file). <serviceDebug includeExceptionDetailInFaults="true" /> If a test causes an unhandled exception on the server the fault is received by the client with a fully populated server stack trace. I can see this by calling the exception's ToString() method. The problem is that this doesn't seem to be output by any of the test runners that I have tried (xUnit, Gallio, MSTest). They appear to just output the Message and the StackTrace properties of the exception. To illustrate what I mean, the following unit test run by MSTest would output three sections: Error Message Error Stack Trace Standard Console Output (contains the information I would like, e.g. "Fault Detail is equal to An ExceptionDetail, likely created by IncludeExceptionDetailInFaults=true, whose value is: ..." try { service.CallMethodWhichCausesException(); } catch (Exception ex) { Console.WriteLine(ex); // this outputs the information I would like throw; } Having this information will make the initial phase of testing and deployment a lot less painful. I know I can just wrap each unit test in a generic exception handler and write the exception to the console and rethrow (as above) within all my unit tests but that seems a very long-winded way of achieving this (and would look pretty awful). Does anyone know if there's any way to get this information included for free whenever an unhandled exception occurs? Is there a setting that I am missing? Is my service configuration lacking in proper fault handling? Perhaps I could write some kind of plug-in / adapter for some unit testing framework? Perhaps theres a different unit testing framework which I should be using instead! My actual set-up is xUnit unit tests executed via Gallio for the development environment, but I do have a separate suite of "smoke tests" written which I would like to be able to have our engineers run via the xUnit GUI test runner (or Gallio or whatever) to simplify the final deployment. Thanks. Adam

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  • VB6 ADO Command to SQL Server

    - by Emtucifor
    I'm getting an inexplicable error with an ADO command in VB6 run against a SQL Server 2005 database. Here's some code to demonstrate the problem: Sub ADOCommand() Dim Conn As ADODB.Connection Dim Rs As ADODB.Recordset Dim Cmd As ADODB.Command Dim ErrorAlertID As Long Dim ErrorTime As Date Set Conn = New ADODB.Connection Conn.ConnectionString = "Provider=SQLOLEDB.1;Integrated Security=SSPI;Initial Catalog=database;Data Source=server" Conn.CursorLocation = adUseClient Conn.Open Set Rs = New ADODB.Recordset Rs.CursorType = adOpenStatic Rs.LockType = adLockReadOnly Set Cmd = New ADODB.Command With Cmd .Prepared = False .CommandText = "ErrorAlertCollect" .CommandType = adCmdStoredProc .NamedParameters = True .Parameters.Append .CreateParameter("@ErrorAlertID", adInteger, adParamOutput) .Parameters.Append .CreateParameter("@CreateTime", adDate, adParamOutput) Set .ActiveConnection = Conn Rs.Open Cmd ErrorAlertID = .Parameters("@ErrorAlertID").Value ErrorTime = .Parameters("@CreateTime").Value End With Debug.Print Rs.State ' Shows 0 - Closed Debug.Print Rs.RecordCount ' Of course this fails since the recordset is closed End Sub So this code was working not too long ago but now it's failing on the last line with the error: Run-time error '3704': Operation is not allowed when the object is closed Why is it closed? I just opened it and the SP returns rows. I ran a trace and this is what the ADO library is actually submitting to the server: declare @p1 int set @p1=1 declare @p2 datetime set @p2=''2010-04-22 15:31:07:770'' exec ErrorAlertCollect @ErrorAlertID=@p1 output,@CreateTime=@p2 output select @p1, @p2 Running this as a separate batch from my query editor yields: Msg 102, Level 15, State 1, Line 4 Incorrect syntax near '2010'. Of course there's an error. Look at the double single quotes in there. What the heck could be causing that? I tried using adDBDate and adDBTime as data types for the date parameter, and they give the same results. When I make the parameters adParamInputOutput, then I get this: declare @p1 int set @p1=default declare @p2 datetime set @p2=default exec ErrorAlertCollect @ErrorAlertID=@p1 output,@CreateTime=@p2 output select @p1, @p2 Running that as a separate batch yields: Msg 156, Level 15, State 1, Line 2 Incorrect syntax near the keyword 'default'. Msg 156, Level 15, State 1, Line 4 Incorrect syntax near the keyword 'default'. What the heck? SQL Server doesn't support this kind of syntax. You can only use the DEFAULT keyword in the actual SP execution statement. I should note that removing the extra single quotes from the above statement makes the SP run fine. ... Oh my. I just figured it out. I guess it's worth posting anyway.

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  • Javascript and the Google Maps API

    - by Tiny Giant Studios
    Hiya coding Ninja's I'm in a spot of bother and my hairline is on the chopping block. When I integrated the maps API on this site, ritaknoetze.com, everything worked perfectly. However, copying that exact code for a different demo website, scarabpaper, the map doesn't show up at all? Could someone show me the ropes on what I'm doing wrong? Here's the code I got from Google itself that I modified for my WordPress theme/installation: JavaScript: <meta name="viewport" content="initial-scale=1.0, user-scalable=no" /> <script type="text/javascript" src="http://maps.google.com/maps/api/js?sensor=false"></script> <script type="text/javascript"> function initialize() { var myLatlng = new google.maps.LatLng(-34.009839, 22.78101); var myOptions = { zoom: 9, center: myLatlng, navigationControl: true, mapTypeControl: false, scaleControl: false, mapTypeId: google.maps.MapTypeId.ROADMAP } var map = new google.maps.Map(document.getElementById("map_canvas"), myOptions); var image = '<?php bloginfo('template_url')?>/assets/googlemaps_marker.png'; var myLatLng = new google.maps.LatLng(-34.009839, 22.78101); var beachMarker = new google.maps.Marker({ position: myLatLng, map: map, icon: image }); } </script> My HTML where the javascript goes: <div class="contact_container"> <div id="map_canvas"></div> <div class="clearfloat"></div> </div> My CSS for the affected divs #map_canvas { width: 880px; height: 300px; margin-left: 10px; margin-bottom: 30px; margin-top: 10px; float: left; border: 1px solid #dedcdc;} .contact_container { /*container for ALL the contact info*/ background-color: #fff; border: 1px solid #dedcdc; width: 900px; margin-top: 30px; padding: 20px; padding-bottom: 0;} Any Help would be greatly appreciated...

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  • .Net Entity Framework SaveChanges is adding without add method

    - by tmfkmoney
    I'm new to the entity framework and I'm really confused about how savechanges works. There's probably a lot of code in my example which could be improved, but here's the problem I'm having. The user enters a bunch of picks. I make sure the user hasn't already entered those picks. Then I add the picks to the database. var db = new myModel() var predictionArray = ticker.Substring(1).Split(','); // Get rid of the initial comma. var user = Membership.GetUser(); var userId = Convert.ToInt32(user.ProviderUserKey); // Get the member with all his predictions for today. var memberQuery = (from member in db.Members where member.user_id == userId select new { member, predictions = from p in member.Predictions where p.start_date == null select p }).First(); // Load all the company ids. foreach (var prediction in memberQuery.predictions) { prediction.CompanyReference.Load(); } var picks = from prediction in predictionArray let data = prediction.Split(':') let companyTicker = data[0] where !(from i in memberQuery.predictions select i.Company.ticker).Contains(companyTicker) select new Prediction { Member = memberQuery.member, Company = db.Companies.Where(c => c.ticker == companyTicker).First(), is_up = data[1] == "up", // This turns up and down into true and false. }; // Save the records to the database. // HERE'S THE PART I DON'T UNDERSTAND. // This saves the records, even though I don't have db.AddToPredictions(pick) foreach (var pick in picks) { db.SaveChanges(); } // This does not save records when the db.SaveChanges outside of a loop of picks. db.SaveChanges(); foreach (var pick in picks) { } // This saves records, but it will insert all the picks exactly once no matter how many picks you have. //The fact you're skipping a pick makes no difference in what gets inserted. var counter = 1; foreach (var pick in picks) { if (counter == 2) { db.SaveChanges(); } counter++; } There's obviously something going on with the context I don't understand. I'm guessing I've somehow loaded my new picks as pending changes, but even if that's true I don't understand I have to loop over them to save changes. Can someone explain this to me?

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  • Is the recent trend toward widescreen (16:9) computer monitors a plus or minus for programmers?

    - by DanM
    It's almost gotten to the point where you can't buy a conventional (4:3) monitor anymore. Pretty much everything is widescreen. This is fine for watching movies or TV, but is it good or bad for programming? My initial thoughts on the issue are that widescreens are a net negative for programmers. Here are some of the disadvantages I see: Poor space utiliziation One disadvantage of widescreens you can't argue with is that they offer poor space utilization for the amount of total pixels you get. For example, my Thinkpad, which I bought just before the widescreen craze, has a 15" monitor with a native resolution of 1600 x 1200. The newer 15.4" Thinkpads run at most 1680 x 1050. So (if you do the math) you get fewer pixels in a wider (but not shorter) package. With desktop monitors, you pay a price in terms of desk space used. Two 1680 x 1050 monitors will simply take up more of your desk than two 1600 x 1200 monitors (assuming equal dot pitch). More scrolling If you compare a 1680 x 1050 monitor to a 1600 x 1200 monitor, you get 80 extra pixels of width but 150 fewer pixels of height. The height reduction means you lose approximately 11 lines of code. That's less you can see on the screen at one time and more scrolling you have to do. This harms productivity, maybe not dramatically, but insidiously. Less room for wide panels Widescreens also mean you lose space for wide but short panels common in programming environments. If you use Visual Studio, for example, your code window will be that much shorter when viewing the Find Results, Task List, or Error List (all of which I use frequently). This isn't to say the 80 pixels of extra width you get with widescreen would never be useful, but I tend to keep my lines of code short, so seeing more lines would be more valuable to me than seeing fewer, longer lines. What do you think? Do you agree/disagree? Are you now using one or more widescreen monitors for development? What resolution are you running on each? Do you ever miss the height of the traditional 4:3 monitor? Would you complain if your monitors were one inch narrower but two inches taller?

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  • WPF Reusing Xaml Effectively

    - by Steve
    Hi, I've recently been working on a project using WPF to produce a diagram. In this I must show text alongside symbols that illustrate information associated with the text. To draw the symbols I initially used some png images I had produced. Within my diagram these images appeared blurry and only looked worse when zoomed in on. To improve on this I decided I would use a vector rather than a rastor image format. Below is the method I used to get the rastor image from a file path: protected Image GetSymbolImage(string symbolPath, int symbolHeight) { Image symbol = new Image(); symbol.Height = symbolHeight; BitmapImage bitmapImage = new BitmapImage(); bitmapImage.BeginInit(); bitmapImage.UriSource = new Uri(symbolPath); bitmapImage.DecodePixelHeight = symbolHeight; bitmapImage.EndInit(); symbol.Source = bitmapImage; return symbol; } Unfortunately this does not recognise vector image formats. So instead I used a method like the following, where "path" is the file path to a vector image of the format .xaml: public static Canvas LoadXamlCanvas(string path) { //if a file exists at the specified path if (File.Exists(path)) { //store the text in the file string text = File.ReadAllText(path); //produce a canvas from the text StringReader stringReader = new StringReader(text); XmlReader xmlReader = XmlReader.Create(stringReader); Canvas c = (Canvas)XamlReader.Load(xmlReader); //return the canvas return c; } return null; } This worked but drastically killed performance when called repeatedly. I found the logic necessary for text to canvas conversion (see above) was the main cause of the performance problem therefore embedding the .xaml images would not alone resolve the performance issue. I tried using this method only on the initial load of my application and storing the resulting canvases in a dictionary that could later be accessed much quicker but I later realised when using the canvases within the dictionary I would have to make copies of them. All the logic I found online associated with making copies used a XamlWriter and XamlReader which would again just introduce a performance problem. The solution I used was to copy the contents of each .xaml image into its own user control and then make use of these user controls where appropriate. This means I now display vector graphics and performance is much better. However this solution to me seems pretty clumsy. I'm new to WPF and wonder if there is some built in way of storing and reusing xaml throughout an application? Apologies for the length of this question. I thought having a record of my attempts might help someone with any similar problem. Thanks.

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  • radgridview delete update in asp.net

    - by abhi
    i have written the follwing to display data from the datagrid and den insert new rows but how do i perform update and delete plss help here's my code using System; using System.Data; using System.Configuration; using System.Collections; using System.Web; using System.Web.Security; using System.Web.UI; using System.Web.UI.WebControls; using System.Web.UI.WebControls.WebParts; using System.Web.UI.HtmlControls; using Telerik.Web.UI; using System.Data.SqlClient; public partial class Default6 : System.Web.UI.Page { string strQry, strCon; SqlDataAdapter da; SqlConnection con; DataSet ds; protected void Page_Load(object sender, EventArgs e) { strCon = "Data Source=MINETDEVDATA; Initial Catalog=ML_SuppliersProd; User Id=sa; Password=@MinetApps7;"; con = new SqlConnection(strCon); strQry = "SELECT * FROM table1"; da = new SqlDataAdapter(strQry, con); SqlCommandBuilder cmdbuild = new SqlCommandBuilder(da); ds = new DataSet(); da.Fill(ds, "table1"); RadGrid1.DataSource = ds.Tables["table1"]; RadGrid1.DataBind(); Label3.Visible = false; Label4.Visible = false; Label5.Visible = false; txtFname.Visible = false; txtLname.Visible = false; txtDesignation.Visible = false; } protected void Submit_Click(object sender, EventArgs e) { Label3.Visible = true; Label4.Visible = true; Label5.Visible = true; txtFname.Visible = true; txtLname.Visible = true; txtDesignation.Visible = true; } protected void Button2_Click(object sender, EventArgs e) { DataSet ds = new DataSet("EmployeeSet"); da.Fill(ds, "table1"); DataTable EmployeeTable = ds.Tables["table1"]; DataRow row = EmployeeTable.NewRow(); row["Fname"] = txtFname.Text.ToString(); row["Lname"] = txtLname.Text.ToString(); row["Designation"] = txtDesignation.Text.ToString(); EmployeeTable.Rows.Add(row); da.Update(ds, "table1"); //RadGrid1.DataSource = ds.Tables["table1"]; //RadGrid1.DataBind(); txtFname.Text = ""; txtLname.Text = ""; txtDesignation.Text = ""; } protected void RadGrid1_DeleteCommand(object source, GridCommandEventArgs e) { } } }

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  • Making a JQuery tooltip retrieve a new value every time the mouse moves.

    - by Micheal Smith
    As i am developing an application that makes use of a tooltip that will display a different value when the user moves the mouse. The user mouses over a table cell and the application then generates a number, the further right the cursor moves in the cell, the higher the value increases. I have created a tooltip that runs and when the cursor mouses over the cell, it does indeed show the correct value. But, when the i move the mouse, it does not show the new value but just the older one. I need to know how to make it update everytime the mouse moves or the value of a variable changes, Any ideas for the problem? <table> <tr id="mon_Section"> <td id="day_Title">Monday</td> <td id="mon_Row"></td> </tr> </table> Below is the document.ready function that calls my function: $(document).ready(function() { $("#mon_Row").mousemove(calculate_Time); }); Below is the function: <script type="text/javascript"> var mon_Pos = 0; var hour = 0; var minute = 0; var orig = 0; var myxpos = 0; function calculate_Time (event) { myxpos = event.pageX; myxpos = myxpos-194; if(myxpos<60) { orig = myxpos; $('#mon_Row').attr("title", orig); } if (myxpos>=60 && myxpos<120) { orig=myxpos; $('#mon_Row').attr("title", orig); } if (myxpos>=120 && myxpos<180) { orig=myxpos; $('#mon_Row').attr("title", orig); Inside the function is the code to generate the tooltip: $('#mon_Row').each(function() { $(this).qtip( { content: { text: false }, position: 'topRight', hide: { fixed: true // Make it fixed so it can be hovered over }, style: { padding: '5px 15px', // Give it some extra padding name: 'dark' // And style it with the preset dark theme } }); }); I know that a new value is being assigned to the cells title attribute because it will display inside the standard small tooltip that a browser will display. The JQuery tooltip will not grab the new value and display it, only the variables initial value when it was called.

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  • iOS: Gesture recogniser for smooth scrolling and flicking a View

    - by AppleDeveloper
    I am building an iPad app where I needed to allow resizing views functionality using divider view provided between two views. This divider view is just a 20px height view between two half screen content views - please refer attached images. When user scrolls this divider view up or down, both content views changes their sizes appropriately. I have extended UIView and implemented this using touchMoved delegate as code given below in touchesMoved delegate. It works fine. The only thing is missing with TouchMoved is you can't flick divider view to top or bottom directly. You have to scroll all the way to top or bottom! To support flicking the view I have tried UIPanGestureRecognizer but I don't see smooth scrolling with it. When I handle split position change in UIGestureRecognizerStateChanged state, just touching divider view flick it to top or bottom. Handling split position change in UIGestureRecognizerStateEnded does the same but I don't see content view resizing with dividerview scrolling! Could someone please tell me how could I achieve both smooth scrolling of divider view with resizing content views(like touchMoved) and flicking the view. Any alternative approach would also fine. Thanks. - (void)touchesMoved:(NSSet *)touches withEvent:(UIEvent *)event { UITouch *touch = [touches anyObject]; if (touch) { CGPoint lastPt = [touch previousLocationInView:self]; CGPoint pt = [touch locationInView:self]; float offset = pt.y - lastPt.y; self.parentViewController.splitPosition = self.parentViewController.splitPosition + offset; } } - (void)handlePan:(UIPanGestureRecognizer*)recognizer { CGPoint translation = [recognizer translationInView:recognizer.view]; CGPoint velocity = [recognizer velocityInView:recognizer.view]; if (recognizer.state == UIGestureRecognizerStateBegan) { } else if (recognizer.state == UIGestureRecognizerStateChanged) { // If I change split position here, I don't see smooth scrolling dividerview...it directly jumps to the top or bottom! self.parentViewController.splitPosition = self.parentViewController.splitPosition + translation.y; } else if (recognizer.state == UIGestureRecognizerStateEnded) { // If I change split position here, the same thing happens at end and I don't see my divider view moving with my scrolling and resizing my views. self.parentViewController.splitPosition = self.parentViewController.splitPosition + translation.y; } } Initial screen Increased top view size by scrolling divider view Top view is totally hidden here but I have to scroll divider view all the way to top. I want to flick the divider view so that it directly goes from any position to top

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  • Calculating rotation and translation matrices between two odometry positions for monocular linear triangulation

    - by user1298891
    Recently I've been trying to implement a system to identify and triangulate the 3D position of an object in a robotic system. The general outline of the process goes as follows: Identify the object using SURF matching, from a set of "training" images to the actual live feed from the camera Move/rotate the robot a certain amount Identify the object using SURF again in this new view Now I have: a set of corresponding 2D points (same object from the two different views), two odometry locations (position + orientation), and camera intrinsics (focal length, principal point, etc.) since it's been calibrated beforehand, so I should be able to create the 2 projection matrices and triangulate using a basic linear triangulation method as in Hartley & Zissermann's book Multiple View Geometry, pg. 312. Solve the AX = 0 equation for each of the corresponding 2D points, then take the average In practice, the triangulation only works when there's almost no change in rotation; if the robot even rotates a slight bit while moving (due to e.g. wheel slippage) then the estimate is way off. This also applies for simulation. Since I can only post two hyperlinks, here's a link to a page with images from the simulation (on the map, the red square is simulated robot position and orientation, and the yellow square is estimated position of the object using linear triangulation.) So you can see that the estimate is thrown way off even by a little rotation, as in Position 2 on that page (that was 15 degrees; if I rotate it any more then the estimate is completely off the map), even in a simulated environment where a perfect calibration matrix is known. In a real environment when I actually move around with the robot, it's worse. There aren't any problems with obtaining point correspondences, nor with actually solving the AX = 0 equation once I compute the A matrix, so I figure it probably has to do with how I'm setting up the two camera projection matrices, specifically how I'm calculating the translation and rotation matrices from the position/orientation info I have relative to the world frame. How I'm doing that right now is: Rotation matrix is composed by creating a 1x3 matrix [0, (change in orientation angle), 0] and then converting that to a 3x3 one using OpenCV's Rodrigues function Translation matrix is composed by rotating the two points (start angle) degrees and then subtracting the final position from the initial position, in order to get the robot's straight and lateral movement relative to its starting orientation Which results in the first projection matrix being K [I | 0] and the second being K [R | T], with R and T calculated as described above. Is there anything I'm doing really wrong here? Or could it possibly be some other problem? Any help would be greatly appreciated.

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