Search Results

Search found 14008 results on 561 pages for 'easy marks'.

Page 520/561 | < Previous Page | 516 517 518 519 520 521 522 523 524 525 526 527  | Next Page >

  • C strange array behaviour

    - by LukeN
    After learning that both strncmp is not what it seems to be and strlcpy not being available on my operating system (Linux), I figured I could try and write it myself. I found a quote from Ulrich Drepper, the libc maintainer, who posted an alternative to strlcpy using mempcpy. I don't have mempcpy either, but it's behaviour was easy to replicate. First of, this is the testcase I have #include <stdio.h> #include <string.h> #define BSIZE 10 void insp(const char* s, int n) { int i; for (i = 0; i < n; i++) printf("%c ", s[i]); printf("\n"); for (i = 0; i < n; i++) printf("%02X ", s[i]); printf("\n"); return; } int copy_string(char *dest, const char *src, int n) { int r = strlen(memcpy(dest, src, n-1)); dest[r] = 0; return r; } int main() { char b[BSIZE]; memset(b, 0, BSIZE); printf("Buffer size is %d", BSIZE); insp(b, BSIZE); printf("\nFirst copy:\n"); copy_string(b, "First", BSIZE); insp(b, BSIZE); printf("b = '%s'\n", b); printf("\nSecond copy:\n"); copy_string(b, "Second", BSIZE); insp(b, BSIZE); printf("b = '%s'\n", b); return 0; } And this is its result: Buffer size is 10 00 00 00 00 00 00 00 00 00 00 First copy: F i r s t b = 46 69 72 73 74 00 62 20 3D 00 b = 'First' Second copy: S e c o n d 53 65 63 6F 6E 64 00 00 01 00 b = 'Second' You can see in the internal representation (the lines insp() created) that there's some noise mixed in, like the printf() format string in the inspection after the first copy, and a foreign 0x01 in the second copy. The strings are copied intact and it correctly handles too long source strings (let's ignore the possible issue with passing 0 as length to copy_string for now, I'll fix that later). But why are there foreign array contents (from the format string) inside my destination? It's as if the destination was actually RESIZED to match the new length.

    Read the article

  • How to use Mozilla ActiveX Control without registry

    - by Andrew McKinlay
    I've been using the IE Browser component that is part of Windows. But I'm running into problems with security settings. For example, users get security warnings on pages with Javascript. So I'm looking at using the Mozilla ActiveX control instead. It's especially nice because it has a compatible interface. It works well if I let it install the control in the registry. But my users don't always have administrator rights to install things in the registry. So I'm trying to figure out how to use the control without registry changes. I'm using DllGetClassObject to get the class factory (IID_ICLASSFACTORY) and then CoRegisterClassObject to register it. All the API calls appear to succeed. And when I create an AtlAxWin window with the CLSID, it also appears to work. But when I try to call Navigate on the AtlAxGetControl it doesn't work - the interface doesn't have Navigate. I would show the code but it's in an obscure language (Suneido) so it wouldn't mean much. An example in C or C++ would be easy for me to translate. Or an example in another dynamic language like Python or Ruby might be helpful. Obviously I'm doing something wrong. Maybe I'm passing the wrong thing to CoRegisterClassObject? The MSDN documentation isn't very clear on what to pass and I haven't found any good examples. Or if there is another approach, I'm ok with that too. Note: I'm using the AtlAxWin window class so I'm not directly creating the control and can't use this approach. Another option is registry free com with a manifest. But again, I couldn't find a good example, especially since I'm not using Visual Studio. I tried to use the MT manifest tool, but couldn't figure it out. I don't think I can use DLL redirection since that doesn't get around the registry issue AFAIK. Another possibility is using WebKit but it seems even harder to use.

    Read the article

  • Route Angular to New Controller after Login

    - by MizAkita
    I'm kind of stuck on how to route my angular app to a new controller after login. I have a simple app, that uses 'loginservice'... after logging in, it then routes to /home which has a different template from the index.html(login page). I want to use /home as the route that displays the partial views of my flightforms controllers. What is the best way to configure my routes so that after login, /home is the default and the routes are called into that particular templates view. Seems easy but I keep getting the /login page when i click on a link which is suppose to pass the partial view into the default.html template: var app= angular.module('myApp', ['ngRoute']); app.config(['$routeProvider', function($routeProvider) { $routeProvider.when('/login', { templateUrl: 'partials/login.html', controller: 'loginCtrl' }); $routeProvider.when('/home', { templateUrl: 'partials/default.html', controller: 'defaultCtrl' }); }]); flightforms.config(['$routeProvider', function($routeProvider){ //sub pages $routeProvider.when('/home', { templateUrl: 'partials/default.html', controller: 'defaultCtrl' }); $routeProvider.when('/status', { templateUrl: 'partials/subpages/home.html', controller: 'statusCtrl' }); $routeProvider.when('/observer-ao', { templateUrl: 'partials/subpages/aobsrv.html', controller: 'obsvaoCtrl' }); $routeProvider.when('/dispatch', { templateUrl: 'partials/subpages/disp.html', controller: 'dispatchCtrl' }); $routeProvider.when('/fieldmgr', { templateUrl: 'partials/subpages/fieldopmgr.html', controller: 'fieldmgrCtrl' }); $routeProvider.when('/obs-backoffice', { templateUrl: 'partials/subpages/obsbkoff.html', controller: 'obsbkoffCtrl' }); $routeProvider.when('/add-user', { templateUrl: 'partials/subpages/users.html', controller: 'userCtrl' }); $routeProvider.otherwise({ redirectTo: '/status' }); }]); app.run(function($rootScope, $location, loginService) { var routespermission=['/home']; //route that require login $rootScope.$on('$routeChangeStart', function(){ if( routespermission.indexOf($location.path()) !=-1) { var connected=loginService.islogged(); connected.then(function(msg) { if(!msg.data) $location.path('/login'); }); } }); }); and my controllers are simple. Here's a sample of what they look like: var flightformsControllers = angular.module('flightformsController', []); flightforms.controller('fieldmgrCtrl', ['$scope','$http','loginService', function($scope,loginService) { $scope.txt='You are logged in'; $scope.logout=function(){ loginService.logout(); } }]); Any ideas on how to get my partials to display in the /home default.html template would be appreciated.

    Read the article

  • Creating multiple heads in remote repository

    - by Jab
    We are looking to move our team (~10 developers) from SVN to mercurial. We are trying to figure out how to manage our workflow. In particular, we are trying to see if creating remote heads is the right solution. We currently have a very large repository with multiple, related projects. They share a lot of code, but pieces of the project are deployed by different teams (3 teams) independent of other portions of the code-base. So each team is working on concurrent large features. The way we currently handles this in SVN are branches. Team1 has a branch for Feature1, same deal for the other teams. When Team1 finishes their change, it gets merged into the trunk and deployed out. The other teams follow suite when their project is complete, merging of course. So my initial thought are using Named Branches for these situations. Team1 makes a Feature1 branch off of the default branch in Hg. Now, here is the question. Should the team PUSH that branch, in it's current/half-state to the repository. This will create a second head in the core repo. My initial reaction was "NO!" as it seems like a bad idea. Handling multiple heads on our repository just sounds awful, but there are some advantages... First, the teams want to setup Continuous Integration to build this branch during their development cycle(months long). This will only work if the CI can pull this branch from the repo. This is something we do now with SVN, copy a CI build and change the branch. Easy. Second, it makes it easier for any team member to jump onto the branch and start working. Without pushing to the core repo, they would have to receive a push from a developer on that team with the changeset information. It is also possible to lose local commits to hardware failure. The chances increase a lot if it's a branch by a single developer who has followed the "don't push until finished" approach. And lastly is just for ease of use. The developers can easily just commit and push on their branch at any time without consequence(as they do today, in their SVN branches). Is there a better way to handle this scenario that I may be missing? I just want a veteran's opinion before moving forward with the strategy. For bug fixes we like the general workflow of mecurial, anonymous branches that only consist of 1-2 commits. The simplicity is great for those cases. By the way, I've read this , great article which seems to favor Named branches.

    Read the article

  • drupal (CMS) or codeigniter (MVC) for creating a new web application?

    - by ajsie
    im going to create a new web application that is very customized. it will contain images, that are fully searchable - in a very, very customized way. when you click on the pictures you can add comments and so on. it requires users to be registered, but the registration/login process will be highly customized too. at the moment im using CodeIgniter for this. But i've read a lot of posts about CMS like Drupal and it sounds like i could let it handle basic stuff, maybe design and other front end work. i have no experience with CMS, in fact, i just started to use a MVC framework like CI and was impressed of how much easier it gets to start developing. so i wonder, if i'm going to create this kind of application, could i use drupal and then add the usual stuff, as i was going to do with CodeIgniter, like controllers, views, models, config files, my own libraries and so on? how does it work on a system like Drupal. how do you code PHP with it as with any MVC framework. it sounds like it has a lot of modules, i just wonder, if i can use it as a MVC framework but have the benefit of having all these basic stuff and design ready to use? cause then it sounds like the best "library" to provide for a web application from scratch. or is it difficult to create a customized app with it? i guess it has modules like images and users, but then how could i customize these so that every image has tags on it and country information, or have every user subscribing to changes to an image, that email will be sent to users and so on? cause i guess its easy to install a module. the question is, how do i customize it. maybe i don't need all that table columns. maybe i want to add/remove business logic. what are the pros and cons with using Drupal for this? is it even the right way to go? can you make a Stackoverflow with Drupal? Facebook? Twitter? Youtube? assuming that you know php of course. share your thoughts cause im totally new on creating a web application! thanks

    Read the article

  • how to remove empty tags in input xml

    - by SGB
    My java module gets a huge input xml from a mainframe. Unfortunately, the mainframe is unable to skip optional elements when it is not a leaf node, with the result that I get a LOT of empty tags in my input : So, <pre><code><SSN>111111111</SSN> <Employment> <Current> <Address> <line1/> <line2/> <line3/> <city/> <state/> <country/> </Address> <Phone> <phonenumber/> <countryCode/> </Phone> </Current> <Previous> <Address> <line1/> <line2/> <line3/> <city/> <state/> <country/> </Address> <Phone> <phonenumber/> <countryCode/> </Phone> </Previous> </Employment> <MaritalStatus>Single</MaritalStatus> </code></pre> should be <SSN>111111111</SSN> <MaritalStatus>Single</MaritalStatus> I use jaxb to unmarshall the input xml string that the mainframe sends it. Is there a clean/ easy way to remove all the empty group tags, or do I have to do this manuall in the code for each element. I have over 35 elements in my input xml, so I would love to it if jaxb itself had a way of doing this automatically? Thanks, SGB

    Read the article

  • WPF: Improving Performance for Running on Older PCs

    - by Phil Sandler
    So, I'm building a WPF app and did a test deployment today, and found that it performed pretty poorly. I was surprised, as we are really not doing much in the way of visual effects or animations. I deployed on two machines: the fastest and the slowest that will need to run the application (the slowest PC has an Intel Celeron 1.80GHz with 2GB RAM). The application ran pretty well on the faster machine, but was choppy on the slower machine. And when I say "choppy", I mean the cursor jumped even just passing it over any open window of the app that had focus. I opened the Task Manager Performance window, and could see that the CPU usage jumped whenever the app had focus and the cursor was moving over it. If I gave focus to another (e.g. Excel), the CPU usage went back down after a second. This happened on both machines, but the choppiness was only noticeable on the slower machine. I had very limited time to tinker on the deployment machines, so didn't do a lot of detailed testing. The app runs fine on my development machine, but I also see the CPU spiking up to 10% there, just running the cursor over the window. I downloaded the WPF performance tool from MS and have been tinkering with it (on my dev machine). The docs say this about the "Frame Rate" metric in the Perforator tool: For applications without animation, this value should be near 0. The app is not doing any heavy animation, but the frame rate stays near 50 when the cursor is over any window. The screens I tested on have column headers in a grid that "highlight" and buttons that change color and appearance when scrolled over. Even moving the mouse on blank areas of the windows cause the same Frame rate and CPU usage (doesn't seem to be related to these minor animations). (Also, I am unable to figure out how to get anything but the two default tools--Perforator and Visual Profiler--installed into the WPF performance tool. That is probably a separate question). I also have Redgate's profiling tool, but I'm not sure if that can shed any light on rendering performance. So, I realize this is not an easy thing to troubleshoot without specifics or sample code (which I can't post). My questions are: What are some general things to look for (or avoid) in the code to improve performance? What steps can I take using the WPF performance tool to narrow down the problem? Is the PC spec listed above (Intel Celeron 1.80GHz with 2GB RAM) too slow to be running even vanilla WPF applications?

    Read the article

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

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

    Read the article

  • What is the best practice to segment c#.net projects based on a single base project

    - by Anthony
    Honestly, I can't word my question any better without describing it. I have a base project (with all its glory, dlls, resources etc) which is a CMS. I need to use this project as a base for othe custom bake projects. This base project is to be maintained and updated among all custom bake projects. I use subversion (Collabnet and Tortise SVN) I have two questions: 1 - Can I use subversion to share the base project among other projects What I mean here is can I "Checkout" the base project into another "Checked Out" project and have both update and commit seperatley. So, to paint a picture, let's say I am working on a custom project and I modify the core/base prject in some way (which I know will suit the others) can I then commit those changes and upon doing so when I update the base project in the other "Checked out" resources will it pull the changes? In short, I would like not to have to manually deploy updated core files whenever I make changes into each seperate project. 2 - If I create a custom file (let's say an webcontrol or aspx page etc) can I have it compile seperatley from the base project Another tricky one to explain. When I publish my web application it creates DLLs based on the namespaces of projects attached to it. So I may have a number of DLLs including the "Website's" namespace DLL, which could simply be website. I want to be able to make a seperate, custom, control which does not compile into those DLLs as the custom files should not rely on those DLLS to run. Is it as simple to set a seperate namespace for those files like CustomFiles.ProjectName for example? Think of the whole idea as adding modules to the .NET project, I don't want the module's code in any of the core DLLs but I do need for module to be able to access the core dlls. (There is no need for the core project to access the module code as it should be one way only in theory, though I reckon it woould not be possible anyway without using JSON/SOAP or something like that, maybe I am wrong.) I want to create a pluggable environment much like that of Joomla/Wordpress as since PHP generally doesn't have to be compiled first I see this is the reason why all this is possible/easy. The idea is to allow pluggable themes, modules etc etc. (I haven't tried simply adding .NET themes after compile/publish but I am assuming this is possible anyway? OR does the compiler need to reference items in the files?)

    Read the article

  • Upgrading Entity Framework 1.0 to 4.0 to include foreign keys

    - by duthiega
    Currently I've been working with Entity Framework 1.0 which is located under a service façade. Below is one of the save methods I've created to either update or insert the device in question. This currently works but, I can't help feel that its a bit of a hack having to set the referenced properties to null then re-attach them just to get an insert to work. The changedDevice already holds these values, so why do I need to assign them again. So, I thought I'll update the model to EF4. That way I can just directly access the foreign keys. However, on doing this I've found that there doesn't seem to be an easy way to add the foreign keys except by removing the entity from the diagram and re-adding it. I don't want to do this as I've already been through all the entity properties renaming them from the DB column names. Can anyone help? /// <summary> /// Saves the non network device. /// </summary> /// <param name="nonNetworkDeviceDto">The non network device dto.</param> public void SaveNonNetworkDevice(NonNetworkDeviceDto nonNetworkDeviceDto) { using (var context = new AssetNetworkEntities2()) { var changedDevice = TransformationHelper.ConvertNonNetworkDeviceDtoToEntity(nonNetworkDeviceDto); if (!nonNetworkDeviceDto.DeviceId.Equals(-1)) { var originalDevice = context.NonNetworkDevices.Include("Status").Include("NonNetworkType").FirstOrDefault( d => d.DeviceId.Equals(nonNetworkDeviceDto.DeviceId)); context.ApplyAllReferencedPropertyChanges(originalDevice, changedDevice); context.ApplyCurrentValues(originalDevice.EntityKey.EntitySetName, changedDevice); } else { var maxNetworkDevice = context.NonNetworkDevices.OrderBy("it.DeviceId DESC").First(); changedDevice.DeviceId = maxNetworkDevice.DeviceId + 1; var status = changedDevice.Status; var nonNetworkType = changedDevice.NonNetworkType; changedDevice.Status = null; changedDevice.NonNetworkType = null; context.AttachTo("DeviceStatuses", status); if (nonNetworkType != null) { context.AttachTo("NonNetworkTypes", nonNetworkType); } changedDevice.Status = status; changedDevice.NonNetworkType = nonNetworkType; context.AddToNonNetworkDevices(changedDevice); } context.SaveChanges(); } }

    Read the article

  • Little CSS problem with Auto height and nested div's

    - by GeekDrop.com
    So I'm finally learning my way around CSS more and have run into a small problem. I have a container div, with a few divs inside of it, one of them is a bit if text (which can be a random height) and an image that will have a MAX height of 200px. I am using a dotted/colored background behind them that needs to auto expand to the height of whichever is the tallest, either the text or the image. Right now when i use height:auto on the container div it works perfect for the random height text: Example Screenshot But it's only adjusting according to the text's height; if the image is taller than the text, the image overflows the bottom of the background dotted/colored box. Example Screenshot The CSS I'm using currently is this: h1 div#like_detailed { margin: 0; font-size: 1.1em; width: 700px; } #details-image img { border: none; clear: left; float: right; margin: -45px 0 0 0; max-height: 200px; padding: 0 7px 0 10px; } #deets-container { background-color: #FEF; border: #190AE7 1px dotted; height: auto; margin-top: 0; margin-bottom: 30px; padding-top: 10px; padding-right: 10px; padding-left: 10px; padding-bottom: 0; } And the HTML for it is this: <div id="deets-container" class="rounded"> <!-- Button --> <div class="likebtnframe">(some code)</div> <!-- Button --> <div class="tweetbtnframe">(some code)</div> <!-- Button --> <ul id="share"> <li><a name="share">(some code)</a></li> </ul> <!-- Submitted By --> <div class="submitter_detailed"><span class="submitter-color smalltext">(some code)</span> (some code)</div> <!-- Image --> <div id="**details-image**">(some code)</div> <!-- Like / Quote --> <h1 id="**like_detailed**">(some code)</h1> </div> I have a feeling this is pretty easy but I'm running out of time to sort it out on my own. Anyone?

    Read the article

  • Issues in Ajax based applications

    - by Sinuhe
    I'm very interested in developing Ajax based applications. This is, loading almost all of the content of the application via XMLHttpRequest, instead of only some combos and widgets. But if I try to do this form scratch, soon I find some problems without an easy solution. I wonder if there is some framework (both client and server side) to deal with this issues. As far as I know, there isn't (but I've searched mainly in Java world). So I am seriously thinking of doing my own framework, at least for my projects. Therefore, in this question I ask for several things. First, the possible problems of an ajax based development. Then, I'm looking for some framework or utility in order to deal with them. Finally, if there is no framework available, what features must it have. Here are the issues I thought: 1 - JavaScript must be enabled. Security paranoia isn't the only problem: a lot of mobile devices couldn't use the application, too. 2 - Sometimes you need to update more than one DIV (e.g. main content, menu and breadcrumbs). 3 - Unknown response type: when you make an Ajax call, you set the callback function too, usually specifying if expected response is a javascript object or in which DIV put the result. But this fails when you get another type of response: for example when the session has expired and the user must log in again. 4 - Browser's refresh, back and forward buttons can be a real pain. User will expect different behaviors depending on the situation. 5 - When search engines indexes a site, only follow links. Thus, content load by Ajax won't "exist" for who doesn't know about it yet. 6 - Users can ask for open a link in a different window/tab. 7 - Address bar doesn't show the "real" page you are in. So, you can't copy the location and send it to a friend or bookmark the page. 8 - If you want to monetize the site, you can put some advertisings. As you don't refresh entire page and you want to change the ad after some time, you have to refresh only the DIV where the ad is. But this can violate the Terms and Conditions of your ad service. In fact, it can go against AdSense TOS. 9 - When you refresh an entire page, all JavaScript gets "cleaned". But in Ajax calls, all JavaScript objects will remain. 10 - You can't easily change your CSS properties.

    Read the article

  • What common routines do you put in your Program.cs for C#

    - by Rick
    I'm interested in any common routine/procedures/methods that you might use in you Program.cs when creating a .NET project. For instance I commonly use the following code in my desktop applications to allow easy upgrades, single instance execution and friendly and simple reporting of uncaught system application errors. using System; using System.Diagnostics; using System.Threading; using System.Windows.Forms; namespace NameoftheAssembly { internal static class Program { /// <summary> /// The main entry point for the application. Modified to check for another running instance on the same computer and to catch and report any errors not explicitly checked for. /// </summary> [STAThread] private static void Main() { //for upgrading and installing newer versions string[] arguments = Environment.GetCommandLineArgs(); if (arguments.GetUpperBound(0) > 0) { foreach (string argument in arguments) { if (argument.Split('=')[0].ToLower().Equals("/u")) { string guid = argument.Split('=')[1]; string path = Environment.GetFolderPath(Environment.SpecialFolder.System); var si = new ProcessStartInfo(path + "\\msiexec.exe", "/x" + guid); Process.Start(si); Application.Exit(); } } //end of upgrade } else { bool onlyInstance = false; var mutex = new Mutex(true, Application.ProductName, out onlyInstance); if (!onlyInstance) { MessageBox.Show("Another copy of this running"); return; } AppDomain.CurrentDomain.UnhandledException += CurrentDomain_UnhandledException; Application.ThreadException += ApplicationThreadException; Application.EnableVisualStyles(); Application.SetCompatibleTextRenderingDefault(false); Application.Run(new Form1()); } } private static void CurrentDomain_UnhandledException(object sender, UnhandledExceptionEventArgs e) { try { var ex = (Exception) e.ExceptionObject; MessageBox.Show("Whoops! Please contact the developers with the following" + " information:\n\n" + ex.Message + ex.StackTrace, " Fatal Error", MessageBoxButtons.OK, MessageBoxIcon.Stop); } catch (Exception) { //do nothing - Another Exception! Wow not a good thing. } finally { Application.Exit(); } } public static void ApplicationThreadException(object sender, ThreadExceptionEventArgs e) { try { MessageBox.Show("Whoops! Please contact the developers with the following" + " information:\n\n" + e.Exception.Message + e.Exception.StackTrace, " Error", MessageBoxButtons.OK, MessageBoxIcon.Stop); } catch (Exception) { //do nothing - Another Exception! Wow not a good thing. } } } } I find these routines to be very helpful. What methods have you found helpful in Program.cs?

    Read the article

  • Javascript self contained sandbox events and client side stack

    - by amnon
    I'm in the process of moving a JSF heavy web application to a REST and mainly JS module application . I've watched "scalable javascript application architecture" by Nicholas Zakas on yui theater (excellent video) and implemented much of the talk with good success but i have some questions : I found the lecture a little confusing in regards to the relationship between modules and sandboxes , on one had to my understanding modules should not be effected by something happening outside of their sandbox and this is why they publish events via the sandbox (and not via the core as they do access the core for hiding base libary) but each module in the application gets a new sandbox ? , shouldn't the sandbox limit events to the modoules using it ? or should events be published cross page ? e.g. : if i have two editable tables but i want to contain each one in a different sandbox and it's events effect only the modules inside that sandbox something like messabe box per table which is a different module/widget how can i do that with sandbox per module , ofcourse i can prefix the events with the moduleid but that creates coupling that i want to avoid ... and i don't want to package modules toghter as one module per combination as i already have 6-7 modules ? while i can hide the base library for small things like id selector etc.. i would still like to use the base library for module dependencies and resource loading and use something like yui loader or dojo.require so in fact i'm hiding the base library but the modules themself are defined and loaded by the base library ... seems a little strange to me libraries don't return simple js objects but usualy wrap them e.g. : u can do something like $$('.classname').each(.. which cleans the code alot , it makes no sense to wrap the base and then in the module create a dependency for the base library by executing .each but not using those features makes a lot of code written which can be left out ... and implemnting that functionality is very bug prone does anyonen have any experience with building a front side stack of this order ? how easy is it to change a base library and/or have modules from different libraries , using yui datatable but doing form validation with dojo ... ? some what of a combination of 2+4 if u choose to do something like i said and load dojo form validation widgets for inputs via yui loader would that mean dojocore is a module and the form module is dependant on it ? Thanks .

    Read the article

  • CSS selectors : should I minimise my use of the class attribute in the HTML or optimise the speed

    - by Laurent Bourgault-Roy
    As I was working on a small website, I decided to use the PageSpeed extension to check if their was some improvement I could do to make the site load faster. However I was quite surprise when it told me that my use of CSS selector was "inefficient". I was always told that you should keep the usage of the class attribute in the HTML to a minimum, but if I understand correctly what PageSpeed tell me, it's much more efficient for the browser to match directly against a class name. It make sense to me, but it also mean that I need to put more CSS classes in my HTML. It also make my .css file a little harder to read. I usually tend to mark my CSS like this : #mainContent p.productDescription em.priceTag { ... } Which make it easy to read : I know this will affect the main content and that it affect something in a paragraph tag (so I wont start to put all sort of layout code in it) that describe a product and its something that need emphasis. However it seem I should rewrite it as .priceTag { ... } Which remove all context information about the style. And if I want to use differently formatted price tag (for example, one in a list on the sidebar and one in a paragraph), I need to use something like that .paragraphPriceTag { ... } .listPriceTag { ... } Which really annoy me since I seem to duplicate the semantic of the HTML in my classes. And that mean I can't put common style in an unqualified .priceTag { ... } and thus I need to replicate the style in both CSS rule, making it harder to make change. (Altough for that I could use multiple class selector, but IE6 dont support them) I believe making code harder to read for the sake of speed has never been really considered a very good practice . Except where it is critical, of course. This is why people use PHP/Ruby/C# etc. instead of C/assembly to code their site. It's easier to write and debug. So I was wondering if I should stick with few CSS classes and complex selector or if I should go the optimisation route and remove my fancy CSS selectors for the sake of speed? Does PageSpeed make over the top recommandation? On most modern computer, will it even make a difference?

    Read the article

  • User has many computers, computers have many attributes in different tables, best way to JOIN?

    - by krismeld
    I have a table for users: USERS: ID | NAME | ---------------- 1 | JOHN | 2 | STEVE | a table for computers: COMPUTERS: ID | USER_ID | ------------------ 13 | 1 | 14 | 1 | a table for processors: PROCESSORS: ID | NAME | --------------------------- 27 | PROCESSOR TYPE 1 | 28 | PROCESSOR TYPE 2 | and a table for harddrives: HARDDRIVES: ID | NAME | ---------------------------| 35 | HARDDRIVE TYPE 25 | 36 | HARDDRIVE TYPE 90 | Each computer can have many attributes from the different attributes tables (processors, harddrives etc), so I have intersection tables like this, to link the attributes to the computers: COMPUTER_PROCESSORS: C_ID | P_ID | --------------| 13 | 27 | 13 | 28 | 14 | 27 | COMPUTER_HARDDRIVES: C_ID | H_ID | --------------| 13 | 35 | So user JOHN, with id 1 owns computer 13 and 14. Computer 13 has processor 27 and 28, and computer 13 has harddrive 35. Computer 14 has processor 27 and no harddrive. Given a user's id, I would like to retrieve a list of that user's computers with each computers attributes. I have figured out a query that gives me a somewhat of a result: SELECT computers.id, processors.id AS p_id, processors.name AS p_name, harddrives.id AS h_id, harddrives.name AS h_name, FROM computers JOIN computer_processors ON (computer_processors.c_id = computers.id) JOIN processors ON (processors.id = computer_processors.p_id) JOIN computer_harddrives ON (computer_harddrives.c_id = computers.id) JOIN harddrives ON (harddrives.id = computer_harddrives.h_id) WHERE computers.user_id = 1 Result: ID | P_ID | P_NAME | H_ID | H_NAME | ----------------------------------------------------------- 13 | 27 | PROCESSOR TYPE 1 | 35 | HARDDRIVE TYPE 25 | 13 | 28 | PROCESSOR TYPE 2 | 35 | HARDDRIVE TYPE 25 | But this has several problems... Computer 14 doesnt show up, because it has no harddrive. Can I somehow make an OUTER JOIN to make sure that all computers show up, even if there a some attributes they don't have? Computer 13 shows up twice, with the same harddrive listet for both. When more attributes are added to a computer (like 3 blocks of ram), the number of rows returned for that computer gets pretty big, and it makes it had to sort the result out in application code. Can I somehow make a query, that groups the two returned rows together? Or a query that returns NULL in the h_name column in the second row, so that all values returned are unique? EDIT: What I would like to return is something like this: ID | P_ID | P_NAME | H_ID | H_NAME | ----------------------------------------------------------- 13 | 27 | PROCESSOR TYPE 1 | 35 | HARDDRIVE TYPE 25 | 13 | 28 | PROCESSOR TYPE 2 | 35 | NULL | 14 | 27 | PROCESSOR TYPE 1 | NULL | NULL | Or whatever result that make it easy to turn it into an array like this [13] => [P_NAME] => [0] => PROCESSOR TYPE 1 [1] => PROCESSOR TYPE 2 [H_NAME] => [0] => HARDDRIVE TYPE 25 [14] => [P_NAME] => [0] => PROCESSOR TYPE 1

    Read the article

  • Having trouble doing an Update with a Linq to Sql object

    - by Pure.Krome
    Hi folks, i've got a simple linq to sql object. I grab it from the database and change a field then save. No rows have been updated. :( When I check the full Sql code that is sent over the wire, I notice that it does an update to the row, not via the primary key but on all the fields via the where clause. Is this normal? I would have thought that it would be easy to update the field(s) with the where clause linking on the Primary Key, instead of where'ing (is that a word :P) on each field. here's the code... using (MyDatabase db = new MyDatabase()) { var boardPost = (from bp in db.BoardPosts where bp.BoardPostId == boardPostId select bp).SingleOrDefault(); if (boardPost != null && boardPost.BoardPostId > 0) { boardPost.ListId = listId; // This changes the value from 0 to 'x' db.SubmitChanges(); } } and here's some sample sql.. exec sp_executesql N'UPDATE [dbo].[BoardPost] SET [ListId] = @p6 WHERE ([BoardPostId] = @p0) AND .... <snip the other fields>',N'@p0 int,@p1 int,@p2 nvarchar(9),@p3 nvarchar(10),@p4 int,@p5 datetime,@p6 int',@p0=1276,@p1=212787,@p2=N'ttreterte',@p3=N'ttreterte3',@p4=1,@p5='2009-09-25 12:32:12.7200000',@p6=72 Now, i know there's a datetime field in this update .. and when i checked the DB it's value was/is '2009-09-25 12:32:12.720' (less zero's, than above) .. so i'm not sure if that is messing up the where clause condition... but still! should it do a where clause on the PK's .. if anything .. for speed! Yes / no ? UPDATE After reading nitzmahone's reply, I then tried playing around with the optimistic concurrency on some values, and it still didn't work :( So then I started some new stuff ... with the optimistic concurrency happening, it includes a where clause on the field it's trying to update. When that happens, it doesn't work. so.. in the above sql, the where clause looks like this ... WHERE ([BoardPostId] = @p0) AND ([ListId] IS NULL) AND ... <rest snipped>) This doesn't sound right! the value in the DB is null, before i do the update. but when i add the ListId value to the where clause (or more to the point, when L2S add's it because of the optomistic concurrecy), it fails to find/match the row. wtf?

    Read the article

  • Check to see if CallResponder is processing

    - by Travesty3
    I'm using Flash Builder 4.6. As a simple example, say I have the following application: <?xml version="1.0" encoding="utf-8"?> <s:Application xmlns:fx="http://ns.adobe.com/mxml/2009" xmlns:s="library://ns.adobe.com/flex/spark" xmlns:sdk="services.sdk.*"> <fx:Script> <![CDATA[ private function btnGetValue_clickHandler():void { getValueResult.token = sdk.getValue(); } private function getValueResultHandler():void { // ... } ]]> </fx:Script> <fx:Declarations> <sdk:SDK id="sdk" fault="{Alert.show(event.fault.faultString +'\n\n'+ event.fault.faultDetail, 'SDK ERROR');}" showBusyCursor="false"/> <s:CallResponder id="getValueResult" result="getValueResultHandler()"/> </fx:Declarations> <s:Button id="btnGetValue" click="btnGetValue_clickHandler()" label="Get Value" /> </s:Application> So when you click on the button, it calls a PHP file and when it gets a result, it calls getValueResultHandler(). Easy enough. But what if the response from the PHP file takes a second or two and the user clicks the button rapidly? Then the result handler function may not get called every time, since the call responder gets a new token before it received the last response. Is there a standard way of resolving this issue? I came up with the following workaround, and it works fine, but it seems like this issue would be common enough to have a more built-in solution. My workaround is: var getValueResultProcessing:Boolean = false; private function btnGetValue_clickHandler():void { var interval:uint = setInterval(function():void { if (!getValueResultProcessing) { getValueResultProcessing = true; getValueResult.token = sdk.getValue(); clearInterval(interval); } }, 100); getValueResult.token = sdk.getValue(); } private function getValueResultHandler():void { getValueResultProcessing = false; // ... } Any better way of resolving this issue?

    Read the article

  • Casting objects in C# (ASP.Net MVC)

    - by Mortanis
    I'm coming from a background in ColdFusion, and finally moving onto something modern, so please bear with me. I'm running into a problem casting objects. I have two database tables that I'm using as Models - Residential and Commercial. Both of them share the majority of their fields, though each has a few unique fields. I've created another class as a container that contains the sum of all property fields. Query the Residential and Commercial, stuff it into my container, cunningly called Property. This works fine. However, I'm having problems aliasing the fields from Residential/Commercial onto Property. It's quite easy to create a method for each property: fillPropertyByResidential(Residential source) and fillPropertyByCommercial(Commercial source), and alias the variables. That also works fine, but quite obviously will copy a bunch of code - all those fields that are shared between the two main Models. So, I'd like a generic fillPropertyBySource() that takes the object, and detects if it's Residential or Commercial, fills the particular fields of each respective type, then do all the fields in common. Except, I gather in C# that variables created inside an If are only in the scope of the if, so I'm not sure how to do this. public property fillPropertyBySource(object source) { property prop = new property(); if (source is Residential) { Residential o = (Residential)source; //Fill Residential only fields } else if (source is Commercial) { Commercial o = (Commercial)source; //Fill Commercial only fields } //Fill fields shared by both prop.price = (int)o.price; prop.bathrooms = (float)o.bathrooms; return prop; } "o" being a Commercial or Residential only exists within the scope of the if. How do I detect the original type of the source object and take action? Bear with me - the shift from ColdFusion into a modern language is pretty..... difficult. More so since I'm used to procedural code and MVC is a massive paradigm shift. Edit: I should include the error: The name 'o' does not exist in the current context For the aliases of price and bathrooms in the shared area.

    Read the article

  • Not allowing characters after Space. Mysql Insert With PHP

    - by Jake
    Ok so I think this is easy but I dont know (I'm a novice to PHP and MySQL). I have a select that is getting data from a table in the database. I am simply taking whatever options the user selects and putting it into a separate table with a php mysql insert statement. But I am having a problem. When I hit submit, everything is submitted properly except for any select options that have spaces don't submit after the first space. For example if the option was COMPUTER REPAIR, all that would get sent is COMPUTER. I will post code if needed, and any help would be greatly appreciated. Thanks! Ok here is my select code: <?php include("./config.php"); $query="SELECT id,name FROM category_names ORDER BY name"; $result = mysql_query ($query); echo"<div style='overflow:auto;width:100%'><label>Categories (Pick three that describe your business)</label><br/><select name='select1'><option value='0'>Please Select A Category</option>"; // printing the list box select command while($catinfo=mysql_fetch_array($result)){//Array or records stored in $nt echo "<option>$catinfo[name]</option><br/> "; } echo"</select></div>"; ?> And here is my insert code ( Just to let you know its got everything not just the select!) ?php require("./config.php"); $companyname = mysql_real_escape_string(addslashes(trim($_REQUEST['name']))); $phone = mysql_real_escape_string(addslashes($_REQUEST['phone'])); $zipcode = mysql_real_escape_string(addslashes($_REQUEST['zipcode'])); $city = mysql_real_escape_string(addslashes($_REQUEST['city'])); $description = mysql_real_escape_string(addslashes($_REQUEST['description'])); $website = mysql_real_escape_string(addslashes($_REQUEST['website'])); $address = mysql_real_escape_string(addslashes($_REQUEST['address'])); $other = mysql_real_escape_string(addslashes($_REQUEST['other'])); $payment = mysql_real_escape_string(addslashes($_REQUEST['payment'])); $products = mysql_real_escape_string(addslashes($_REQUEST['products'])); $email = mysql_real_escape_string(addslashes($_REQUEST['email'])); $select1 = mysql_real_escape_string(addslashes($_REQUEST['select1'])); $select2 = mysql_real_escape_string(addslashes($_REQUEST['select2'])); $select3 = mysql_real_escape_string(addslashes($_REQUEST['select3'])); $save=$_POST['save']; if(!empty($save)){ $sql="INSERT INTO gj (name, phone, city, zipcode, description, dateadded, website, address1, other2, payment_options, Products, email,cat1,cat2,cat3) VALUES ('$companyname','$phone','$city','$zipcode','$description',curdate(),'$website','$address','$other','$payment','$products','$email','$select1','$select2','$select3')"; if (!mysql_query($sql,$link)) { die('Error: ' . mysql_error()); } echo "<br/><h2><font color='green' style='font-size:15px'>1 business added</font></h2>"; mysql_close($link); } ?>

    Read the article

  • Counting point size based on chart area during zooming/unzoomin

    - by Gacek
    Hi folks. I heave a quite simple task. I know (I suppose) it should be easy, but from the reasons I cannot understand, I try to solve it since 2 days and I don't know where I'm making the mistake. So, the problem is as follows: - we have a chart with some points - The chart starts with some known area and points have known size - we would like to "emulate" the zooming effect. So when we zoom to some part of the chart, the size of points is getting proportionally bigger. In other words, the smaller part of the chart we select, the bigger the point should get. So, we have something like that. We know this two parameters: initialArea; // Initial area - area of the whole chart, counted as width*height initialSize; // initial size of the points Now lets assume we are handling some kind of OnZoom event. We selected some part of the chart and would like to count the current size of the points float CountSizeOnZoom() { float currentArea = CountArea(...); // the area is counted for us. float currentSize = initialSize * initialArea / currentArea; return currentSize; } And it works. But the rate of change is too fast. In other words, the points are getting really big too soon. So I would like the currentSize to be invertly proportional to currentArea, but with some scaling coefficient. So I created the second function: float CountSizeOnZoom() { float currentArea = CountArea(...); % the area is counted for us. // Lets assume we want the size of points to change ten times slower, than area of the chart changed. float currentSize = initialSize + 0.1f* initialSize * ((initialArea / currentArea) -1); return currentSize; } Lets do some calculations in mind. if currentArea is smaller than initialArea, initialArea/currentArea > 1 and then we add "something" small and postive to initialSize. Checked, it works. Lets check what happens if we would un-zoom. currentArea will be equal to initialArea, so we would have 0 at the right side (1-1), so new size should be equal to initialSize. Right? Yeah. So lets check it... and it doesn't work. My question is: where is the mistake? Or maybe you have any ideas how to count this scaled size depending on current area in some other way?

    Read the article

  • Adding Insert Row in tableView

    - by user333624
    Hello everyone, I have a tableView that loads its data directly from a Core Data table with a NSFetchedResultsController. I'm not using an intermediate NSMutableArray to store the objects from the fetch results; I basically implemented inside my UITableViewController the protocol method numberOfRowsInSection and it returns the numberOfObjects inside a NSFetchedResultsSectionInfo. id <NSFetchedResultsSectionInfo> sectionInfo = [[fetchedResultsController sections] objectAtIndex:section]; and then I configure the cell content by implementing configureCell:atIndexPath and retrieving object info from the fetchedResultController but right now I have a generic configuration for any object (to avoid complications) cell.textLabel.text = @"categoria"; I also have a NavigationBar with a custom edit button at the right that loads my own selector called customSetEditing. What I'm trying to accomplish is to load an "Insert Cell" at the beginning of the tableView so when I tap it, it creates a new record. This last part is easy to implement the problem is that I dont's seem to be able to load the insert row or any row when I tap on the navigation bar edit button. this is the code for my customSetEditing: - (void) customSetEditing { [super setEditing:YES animated:YES]; [self.tableView setEditing:YES animated:YES]; [[self tableView] beginUpdates]; //[[self tableView] beginUpdates]; UIBarButtonItem *customDoneButtonItem = [[UIBarButtonItem alloc] initWithBarButtonSystemItem:UIBarButtonSystemItemDone target:self action:@selector(customDone)]; [self.navigationItem.rightBarButtonItem release]; self.navigationItem.rightBarButtonItem = customDoneButtonItem; //[categoriasArray insertObject:[NSNull null] atIndex:0]; NSMutableArray *indexPaths = [[NSMutableArray alloc] initWithObjects:[NSIndexPath indexPathForRow:0 inSection:0],nil ]; [self.tableView insertRowsAtIndexPaths:indexPaths withRowAnimation:UITableViewRowAnimationTop]; //[indexPaths release]; [self.tableView reloadData];} Before adding the:[self.tableView reloadData]; I was getting an out of bounds error plus a program crash and although the program is not crashing it is not loading anything. I have seen many examples of similar situations in stackoverflow (by the way is an excellent forum with very helpful and nice people) none of the examples seems to work for me. Any ideas?

    Read the article

  • B-trees, databases, sequential inputs, and speed.

    - by IanC
    I know from experience that b-trees have awful performance when data is added to them sequentially (regardless of the direction). However, when data is added randomly, best performance is obtained. This is easy to demonstrate with the likes of an RB-Tree. Sequential writes cause a maximum number of tree balances to be performed. I know very few databases use binary trees, but rather used n-order balanced trees. I logically assume they suffer a similar fate to binary trees when it comes to sequential inputs. This sparked my curiosity. If this is so, then one could deduce that writing sequential IDs (such as in IDENTITY(1,1)) would cause multiple re-balances of the tree to occur. I have seen many posts argue against GUIDs as "these will cause random writes". I never use GUIDs, but it struck me that this "bad" point was in fact a good point. So I decided to test it. Here is my code: SET ANSI_NULLS ON GO SET QUOTED_IDENTIFIER ON GO CREATE TABLE [dbo].[T1]( [ID] [int] NOT NULL CONSTRAINT [T1_1] PRIMARY KEY CLUSTERED ([ID] ASC) ) GO CREATE TABLE [dbo].[T2]( [ID] [uniqueidentifier] NOT NULL CONSTRAINT [T2_1] PRIMARY KEY CLUSTERED ([ID] ASC) ) GO declare @i int, @t1 datetime, @t2 datetime, @t3 datetime, @c char(300) set @t1 = GETDATE() set @i = 1 while @i < 2000 begin insert into T2 values (NEWID(), @c) set @i = @i + 1 end set @t2 = GETDATE() WAITFOR delay '0:0:10' set @t3 = GETDATE() set @i = 1 while @i < 2000 begin insert into T1 values (@i, @c) set @i = @i + 1 end select DATEDIFF(ms, @t1, @t2) AS [Int], DATEDIFF(ms, @t3, getdate()) AS [GUID] drop table T1 drop table T2 Note that I am not subtracting any time for the creation of the GUID nor for the considerably extra size of the row. The results on my machine were as follows: Int: 17,340 ms GUID: 6,746 ms This means that in this test, random inserts of 16 bytes was almost 3 times faster than sequential inserts of 4 bytes. Would anyone like to comment on this? Ps. I get that this isn't a question. It's an invite to discussion, and that is relevant to learning optimum programming.

    Read the article

  • Code Golf: Countdown Number Game

    - by Noldorin
    Challenge Here is the task, inspired by the well-known British TV game show Countdown. The challenge should be pretty clear even without any knowledge of the game, but feel free to ask for clarifications. And if you fancy seeing a clip of this game in action, check out this YouTube clip. It features the wonderful late Richard Whitely in 1997. You are given 6 numbers, chosen at random from the set {1, 2, 3, 4, 5, 6, 8, 9, 10, 25, 50, 75, 100}, and a random target number between 100 and 999. The aim is to make use the six given numbers and the four common arithmetic operations (addition, subtraction, multiplication, division; all over the rational numbers) to generate the target - or as close as possible either side. Each number may only be used once at most, while each arithmetic operator may be used any number of times (including zero.) Note that it does not matter how many numbers are used. Write a function that takes the target number and set of 6 numbers (can be represented as list/collection/array/sequence) and returns the solution in any standard numerical notation (e.g. infix, prefix, postfix). The function must always return the closest-possible result to the target, and must run in at most 1 minute on a standard PC. Note that in the case where more than one solution exists, any single solution is sufficient. Examples: {50, 100, 4, 2, 2, 4}, target 203 e.g. 100 * 2 + 2 + (4 / 4) e.g. (100 + 50) * 4 * 2 / (4 + 2) {25, 4, 9, 2, 3, 10}, target 465 e.g. (25 + 10 - 4) * (9 * 2 - 3) {9, 8, 10, 5, 9, 7), target 241 e.g. ((10 + 9) * 9 * 7) + 8) / 5 Rules Other than mentioned in the problem statement, there are no further restrictions. You may write the function in any standard language (standard I/O is not necessary). The aim as always is to solve the task with the smallest number of characters of code. Saying that, I may not simply accept the answer with the shortest code. I'll also be looking at elegance of the code and time complexity of the algorithm! My Solution I'm attempting an F# solution when I find the free time - will post it here when I have something! Format Please post all answers in the following format for the purpose of easy comparison: Language Number of characters: ??? Fully obfuscated function: (code here) Clear (ideally commented) function: (code here) Any notes on the algorithm/clever shortcuts it takes.

    Read the article

  • C++ class member functions instantiated by traits

    - by Jive Dadson
    I am reluctant to say I can't figure this out, but I can't figure this out. I've googled and searched Stack Overflow, and come up empty. The abstract, and possibly overly vague form of the question is, how can I use the traits-pattern to instantiate non-virtual member functions? The question came up while modernizing a set of multivariate function optimizers that I wrote more than 10 years ago. The optimizers all operate by selecting a straight-line path through the parameter space away from the current best point (the "update"), then finding a better point on that line (the "line search"), then testing for the "done" condition, and if not done, iterating. There are different methods for doing the update, the line-search, and conceivably for the done test, and other things. Mix and match. Different update formulae require different state-variable data. For example, the LMQN update requires a vector, and the BFGS update requires a matrix. If evaluating gradients is cheap, the line-search should do so. If not, it should use function evaluations only. Some methods require more accurate line-searches than others. Those are just some examples. The original version instantiates several of the combinations by means of virtual functions. Some traits are selected by setting mode bits that are tested at runtime. Yuck. It would be trivial to define the traits with #define's and the member functions with #ifdef's and macros. But that's so twenty years ago. It bugs me that I cannot figure out a whiz-bang modern way. If there were only one trait that varied, I could use the curiously recurring template pattern. But I see no way to extend that to arbitrary combinations of traits. I tried doing it using boost::enable_if, etc.. The specialized state information was easy. I managed to get the functions done, but only by resorting to non-friend external functions that have the this-pointer as a parameter. I never even figured out how to make the functions friends, much less member functions. The compiler (VC++ 2008) always complained that things didn't match. I would yell, "SFINAE, you moron!" but the moron is probably me. Perhaps tag-dispatch is the key. I haven't gotten very deeply into that. Surely it's possible, right? If so, what is best practice?

    Read the article

< Previous Page | 516 517 518 519 520 521 522 523 524 525 526 527  | Next Page >