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  • ELMAH - Using custom error pages to collecting user feedback

    - by vdh_ant
    Hey guys I'm looking at using ELMAH for the first time but have a requirement that needs to be met that I'm not sure how to go about achieving... Basically, I am going to configure ELMAH to work under asp.net MVC and get it to log errors to the database when they occur. On top of this I be using customErrors to direct the user to a friendly message page when an error occurs. Fairly standard stuff... The requirement is that on this custom error page I have a form which enables to user to provide extra information if they wish. Now the problem arises due to the fact that at this point the error is already logged and I need to associate the loged error with the users feedback. Normally, if I was using my own custom implementation, after I log the error I would pass through the ID of the error to the custom error page so that an association can be made. But because of the way that ELMAH works, I don't think the same is quite possible. Hence I was wondering how people thought that one might go about doing this.... Cheers UPDATE: My solution to the problem is as follows: public class UserCurrentConextUsingWebContext : IUserCurrentConext { private const string _StoredExceptionName = "System.StoredException."; private const string _StoredExceptionIdName = "System.StoredExceptionId."; public virtual string UniqueAddress { get { return HttpContext.Current.Request.UserHostAddress; } } public Exception StoredException { get { return HttpContext.Current.Application[_StoredExceptionName + this.UniqueAddress] as Exception; } set { HttpContext.Current.Application[_StoredExceptionName + this.UniqueAddress] = value; } } public string StoredExceptionId { get { return HttpContext.Current.Application[_StoredExceptionIdName + this.UniqueAddress] as string; } set { HttpContext.Current.Application[_StoredExceptionIdName + this.UniqueAddress] = value; } } } Then when the error occurs, I have something like this in my Global.asax: public void ErrorLog_Logged(object sender, ErrorLoggedEventArgs args) { var item = new UserCurrentConextUsingWebContext(); item.StoredException = args.Entry.Error.Exception; item.StoredExceptionId = args.Entry.Id; } Then where ever you are later you can pull out the details by var item = new UserCurrentConextUsingWebContext(); var error = item.StoredException; var errorId = item.StoredExceptionId; item.StoredException = null; item.StoredExceptionId = null; Note this isn't 100% perfect as its possible for the same IP to have multiple requests to have errors at the same time. But the likely hood of that happening is remote. And this solution is independent of the session, which in our case is important, also some errors can cause sessions to be terminated, etc. Hence why this approach has worked nicely for us.

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  • Learning MVC - Maintaining model state

    - by GenericTypeTea
    First of all, I'm very new to MVC. Bought the books, but not got the T-Shirt yet. I've put together my first little application, but I'm looking at the way I'm maintaining my model and I don't think it looks right. My form contains the following: <% using (Html.BeginForm("Reconfigured", null, FormMethod.Post, new { id = "configurationForm" })) { %> <%= Html.DropDownList("selectedCompany", new SelectList(Model.Companies, Model.SelectedCompany), new { onchange = "$('#configurationForm').submit()" })%> <%= Html.DropDownList("selectedDepartment", new SelectList(Model.Departments, Model.SelectedDepartment), new { onchange = "$('#configurationForm').submit()" })%> <%=Html.TextArea("comment", Model.Comment) %> <%} %> My controller has the following: public ActionResult Index(string company, string department, string comment) { TestModel form = new TestModel(); form.Departments = _someRepository.GetList(); form.Companies = _someRepository.GetList(); form.Comment = comment; form.SelectedCompany = company; form.SelectedDepartment = department; return View(form); } [HttpPost] public ActionResult Reconfigured(string selectedCompany, string selectedDepartment, string comment) { return RedirectToAction("Index", new { company = selectedCompany, department = selectedDepartment, comment = comment}); } And finally, this is my route: routes.MapRoute( "Default", "{controller}/{company}/{department}", new { controller = "CompanyController", action = "Index", company="", department="" } ); Now, every time I change DropDownList value, all my values are maintained. I end up with a URL like the following after the Reconfigure action is called: http://localhost/Main/Index/Company/Sales?comment=Foo%20Bar Ideally I'd like the URL to remain as: http://localhost/Main/Index My routing object is probably wrong. This can't be the right way? It seems totally wrong to me as for each extra field I add, I have to add the property into the Index() method? I had a look at this answer where the form is passed through TempData. This is obviously an improvement, but it's not strongly typed? Is there a way to do something similar but have it strongly typed? This may be a simple-enough question, but the curse of 10 years of WinForms/WebForms makes this MVC malarky hard to get your head 'round.

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  • Help to understand and recode javascript function to deal with special characters.

    - by Cesar Lopez
    Hi all, I am trying to rewrite a javascript function since I was told this function its a bit nasty peace of code and it could be nicely written by a very kind user from here. I have been trying to understand what the function does, therefore I could rewrite it properly, but since I dont fully understand how it works its a very difficult task. Therefore I am looking for help and directions (NOT THE SOLUTION AS I WANT TO LEARN MYSELF) to understand and rewrite this function in a nicer way. The function its been made for dealing with special characters, and I know that it loops through the string sent to it, search for special characters, and add what it needs to the string to make it a valid string. I have been trying to use value.replace(/"/gi,"/""), but surely I am doing it wrong as it crashes. Could anybody tell me where to start to recode function? Any help would be appreciated. My comments on the function are in capital letters. Code <script type="text/javascript"> function convertString(value){ for(var z=0; z <= value.length -1; z++) { //if current character is a backslash||WHY IS IT CHECKING FOR \\,\\r\\n,and \\n? if(value.substring(z, z + 1)=="\\" && (value.substring(z, z + 4)!="\\r\\n" && value.substring(z, z + 2)!="\\n")) {//WHY IS IT ADDING \\\\ TO THE STRING? value = value.substring(0, z) + "\\\\" + value.substring(z + 1, value.length); z++; } if(value.substring(z, z + 1)=="\\" && value.substring(z, z + 4)=="\\r\\n") {//WHY IS IT ADDING 4 TO Z IN THIS CASE? z = z+4; } if(value.substring(z, z + 1)=="\\" && value.substring(z, z + 2)=="\\n") {//WHY IS IT ADDING 2 TO Z IN THIS CASE? z = z+2; } } //replace " with \" //loop through each character for(var x = 0; x <= value.length -1; x++){ //if current character is a quote if(value.substring(x, x + 1)=="\""){//THIS IS TO FIND \, BUT HAVENT THIS BEEN DONE BEFFORE? //concatenate: value up to the quote + \" + value AFTER the quote||WHY IS IT ADDING \\ BEFORE \"? value = value.substring(0, x) + "\\\"" + value.substring(x + 1, value.length); //account for extra character x++; } } //return the modified string return(value); } <script> Comments within the code on capital letters are my questions about the function as I mention above. I would appreciate any help, orientation, advise, BUT NOT THE SOLUTION PLEASE AS I DO WANT TO LEARN.

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  • Vim + OmniCppComplete: Completing on Class Members which are STL containers

    - by Robert S. Barnes
    Completion on class members which are STL containers is failing. Completion on local objects which are STL containers works fine. For example, given the following files: // foo.h #include <string> class foo { public: void set_str(const std::string &); std::string get_str_reverse( void ); private: std::string str; }; // foo.cpp #include "foo.h" using std::string; string foo::get_str_reverse ( void ) { string temp; temp.assign(str); reverse(temp.begin(), temp.end()); return temp; } /* ----- end of method foo::get_str ----- */ void foo::set_str ( const string &s ) { str.assign(s); } /* ----- end of method foo::set_str ----- */ I've generated the tags for these two files using: ctags -R --c++-kinds=+pl --fields=+iaS --extra=+q . When I type temp. in the cpp I get a list of string member functions as expected. But if I type str. omnicppcomplete spits out "Pattern Not Found". I've noticed that the temp. completion only works if I have the using std::string; declaration. How do I get completion to work on my class members which are STL containers? Edit I found that completion on members which are STL containers works if I make the follow modifications to the header: // foo.h #include <string> using std::string; class foo { public: void set_str(const string &); string get_str_reverse( void ); private: string str; }; Basically, if I add using std::string; and then remove the std:: name space qualifier from the string str; member and regenerate the tags file then OmniCppComplete is able to do completion on str.. It doesn't seem to matter whether or not I have let OmniCpp_DefaultNamespaces = ["std", "_GLIBCXX_STD"] set in the .vimrc. The problem is that putting using declarations in header files seems like a big no-no, so I'm back to square one.

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  • How can I limit the cache used by copying so there is still memory available for other cache?

    - by Peter
    Basic situation: I am copying some NTFS disks in openSuSE. Each one is 2TB. When I do this, the system runs slow. My guesses: I believe it is likely due to caching. Linux decides to discard useful cache (eg. kde4 bloat, virtual machine disks, LibreOffice binaries, Thunderbird binaries, etc.) and instead fill all available memory (24 GB total) with stuff from the copying disks, which will be read only once, then written and never used again. So then any time I use these apps (or kde4), the disk needs to be read again, and reading the bloat off the disk again makes things freeze/hiccup. Due to the cache being gone and the fact that these bloated applications need lots of cache, this makes the system horribly slow. Since it is USB,the disk and disk controller are not the bottleneck, so using ionice does not make it faster. I believe it is the cache rather than just the motherboard going too slow, because if I stop everything copying, it still runs choppy for a while until it recaches everything. And if I restart the copying, it takes a minute before it is choppy again. But also, I can limit it to around 40 MB/s, and it runs faster again (not because it has the right things cached, but because the motherboard busses have lots of extra bandwidth for the system disks). I can fully accept a performance loss from my motherboard's IO capability being completely consumed (which is 100% used, meaning 0% wasted power which makes me happy), but I can't accept that this caching mechanism performs so terribly in this specific use case. # free total used free shared buffers cached Mem: 24731556 24531876 199680 0 8834056 12998916 -/+ buffers/cache: 2698904 22032652 Swap: 4194300 24764 4169536 I also tried the same thing on Ubuntu, which causes a total system hang instead. ;) And to clarify, I am not asking how to leave memory free for the "system", but for "cache". I know that cache memory is automatically given back to the system when needed, but my problem is that it is not reserved for caching of specific things. Question: Is there some way to tell these copy operations to limit memory usage so some important things remain cached, and therefore any slowdowns are a result of normal disk usage and not rereading the same commonly used files? For example, is there a setting of max memory per process/user/file system allowed to be used as cache/buffers?

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  • Arguments for moving from LINQtoSQL to Nhibernate?

    - by sah302
    Backstory: Hi all, I just spent a lot of time reading many of the LINQ vs Nhibernate threads here and on other sites. I work in a small development team of 4 people and we don't even have really any super experienced developers. We work for a small company that has a lot of technical needs but not enough developers to implement them (and hiring more is out of the question right now). Typically our projects (which individually are fairly small) have been coded separately and weren't really layered in anyway, code wasn't re-used, no class libraries, and we just use the LINQtoSQL .dbml files for our pojects, we really don't even use objects but pass around values and stuff, the only time we use objects is when inserting to a database (heck not even querying since you don't need to assign it to a type and can just bind to gridview). Despite all this as I said our company has a lot of technical needs, no one could come to us for a year and we would have plenty of work to implement requested features. Well I have decided to change that a bit first by creating class libraries and actually adding layers to our applications. I am trying to meet these guys halfway by still using LINQtoSQL as the ORM yet and still use VB as the language. However I am finding it a b***h of a time dealing with so many thing in LINQtoSQL that I found easy in Nhibernate (automatic handling of the session, criteria creation easier than expression trees, generic an dynamic querying easier etc.) So... Question: How can I convince my lead developers and other senior programmers that switching to Nhibernate is a good thing? That being in control of our domain objects is a good thing? That being able to implement interfaces is a good? I've tried exlpaining the advantages of this before but it's not understood by them because they've never programmed in a true OO & layered way. Also one of the counter arguments to this I can see is sqlMetal generates those classes automatically and therefore it saves a lot of time. I can't really counter that other than saying spending more time on infrastructure to make it more scalable and flexible is good, but they can't see how. Again, I know the features and advantages (somewhat enough I believe) of each, but I need arguments applicable to my context, hence why I provided the context. I just am not a very good arguer I guess. (Caveat: For all the LINQtoSQL lovers, I may just not be super proficient as LINQ, but I find it very cumbersome that you are required to download some extra library for dynamic queries which don't by default support guid comparisons, and I also find the way of updating entitites to be cumbersome as well in terms of data context managing, so it could just be that I suck hehe.)

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  • Mocking a concrete class : templates and avoiding conditional compilation

    - by AshirusNW
    I'm trying to testing a concrete object with this sort of structure. class Database { public: Database(Server server) : server_(server) {} int Query(const char* expression) { server_.Connect(); return server_.ExecuteQuery(); } private: Server server_; }; i.e. it has no virtual functions, let alone a well-defined interface. I want to a fake database which calls mock services for testing. Even worse, I want the same code to be either built against the real version or the fake so that the same testing code can both: Test the real Database implementation - for integration tests Test the fake implementation, which calls mock services To solve this, I'm using a templated fake, like this: #ifndef INTEGRATION_TESTS class FakeDatabase { public: FakeDatabase() : realDb_(mockServer_) {} int Query(const char* expression) { MOCK_EXPECT_CALL(mockServer_, Query, 3); return realDb_.Query(); } private: // in non-INTEGRATION_TESTS builds, Server is a mock Server with // extra testing methods that allows mocking Server mockServer_; Database realDb_; }; #endif template <class T> class TestDatabaseContainer { public: int Query(const char* expression) { int result = database_.Query(expression); std::cout << "LOG: " << result << endl; return result; } private: T database_; }; Edit: Note the fake Database must call the real Database (but with a mock Server). Now to switch between them I'm planning the following test framework: class DatabaseTests { public: #ifdef INTEGRATION_TESTS typedef TestDatabaseContainer<Database> TestDatabase ; #else typedef TestDatabaseContainer<FakeDatabase> TestDatabase ; #endif TestDatabase& GetDb() { return _testDatabase; } private: TestDatabase _testDatabase; }; class QueryTestCase : public DatabaseTests { public: void TestStep1() { ASSERT(GetDb().Query(static_cast<const char *>("")) == 3); return; } }; I'm not a big fan of that compile-time switching between the real and the fake. So, my question is: Whether there's a better way of switching between Database and FakeDatabase? For instance, is it possible to do it at runtime in a clean fashion? I like to avoid #ifdefs. Also, if anyone has a better way of making a fake class that mimics a concrete class, I'd appreciate it. I don't want to have templated code all over the actual test code (QueryTestCase class). Feel free to critique the code style itself, too. You can see a compiled version of this code on codepad.

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  • Which MarkDown (WMD) javascript editor should I use?

    - by Edan Maor
    Background I'm working on an application which requires user-entered content, and I've decided to use a StackOverflow-style MarkDown editor. After researching this topic for the last few days, I realize there are numerous forks of the base WMD editor, some with a few basic enhancements and some with serious differences from the StackOverflow one. Since this will be the heart of the application, I'd like to start with the best code base I can. I'd be happy if anyone can recommend which one of the many solutions out there best fits my needs. Below is requirements, plus what I've managed to find already. I'm hoping this question will help me decide which version to go with, and maybe help me discover a port out there that's an even better fit for my needs. The requirements for my project Live Preview Multiple editors on the same page (not know how many in advance, since the user can dynamically add another editing box). Ability to extend with extra buttons (I'd like a button to upload a picture, instead of just adding an img url). Ability to dynamically show/hide the edit box (and only see the preview box). Not an absolute must, but I'd prefer to stick as close to StackOverflow's look and feel, since it's well known. Don't know if this matters, but the backend is written in Django. Editors I've looked at Here are a few of the code bases I've looked at, with thoughts. Obviously, I might be missing another solution out there. The derobins version. From what I can tell, this is the official StackOverflow version. Seems like it doesn't support multiple editors on one page. JQuery.MarkEdit. Looks very good, but is pretty different from the StackOverflow version. MooWMD. Looks like the winner right now, but I'm a little concerned since it looks less active/hackable than MarkEdit. The wmd-new version. Not sure, looks like an old codebase without much use. The SocialSite branch. Seems like it's not for public use.

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  • do I need to close an audio Clip?

    - by Michael
    have an application that processes real-time data and is supposed to beep when a certain event occurs. The triggering event can occur multiple times per second, and if the beep is already playing when another event triggers the code is just supposed to ignore it (as opposed to interrupting the current beep and starting a new one). Here is the basic code: Clip clickClip public void prepareProcess() { super.prepareProcess(); clickClip = null; try { clipFile = new File("C:/WINDOWS/Media/CHIMES.wav"); ais = AudioSystem.getAudioInputStream(clipFile); clickClip = AudioSystem.getClip(); clickClip.open(ais); fileIsLoaded = true; } catch (Exception ex) { clickClip = null; fileIsLoaded = false; } } public void playSound() { if (fileIsLoaded) { if ((clickClip==null) || (!clickClip.isRunning())) { try { clickClip.setFramePosition(0); clickClip.start(); } catch (Exception ex) { System.out.println("Cannot play click noise"); ex.printStackTrace(); } } } The prepareProcess method gets run once in the beginning, and the playSound method is called every time a triggering event occurs. My question is: do I need to close the clickClip object? I know I could add an actionListener to monitor for a Stop event, but since the event occurs so frequently I'm worried the extra processing is going to slow down the real-time data collection. The code seems to run fine, but my worry is memory leaks. The code above is based on an example I found while searching the net, but the example used an actionListener to close the Clip specifically "to eliminate memory leaks that would occur when the stop method wasn't implemented". My program is intended to run for hours so any memory leaks I have will cause problems. I'll be honest: I have no idea how to verify whether or not I've got a problem. I'm using Netbeans, and running the memory profiler just gave me a huge list of things that I don't know how to read. This is supposed to be the simple part of the program, and I'm spending hours on it. Any help would be greatly appreciated! Michael

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  • Various GPS Android Functionality Questions..

    - by Tyler
    Hello - I have a few questions (so far) with the the LocationManager on Android and GPS in general.. Feel free to answer any number of the questions below, and I appreciate your help in advance! (I noticed this stuff doesn't appear to be documented very well, so hopefully these questions will help others out too!) 1) I am using the following code, but I think there may be extra fluff in here that I do not need. Can you tell me if I can delete any of this? LocationManager lm = (LocationManager) getSystemService(Context.LOCATION_SERVICE); LocationListener locationListener = new MyLocationListener(); lm.requestLocationUpdates(LocationManager.GPS_PROVIDER, 0, 0, locationListener); LocationProvider locationProvider = lm.getProvider("gps"); Location currentLocation = lm.getLastKnownLocation(locationProvider.getName()); 2) Is there a way to hold off on the last step (accessing "getLastKnownLocation" until after I am sure I have a GPS lock? What happens if this is called and GPS is still looking for signal? 3) MOST importantly, I want to ensure I have a GPS lock before I proceed to my next method, so is there a way to check to see if GPS is locked on and getLastKnownLocation is up to date? 4) Is there a way to 'shut down' the GPS listener once it does receive a lock and getLastKnownLocation is updated? I don't see a need to keep this running for my application once I have obtained a lock.. 5) Can you please confirm my assumption that "getLastKnownLocation" is updated frequently as the receiver moves? 6) In my code, I also have a class called "MyLocationListener" (code below) that I honestly just took from another example.. Is this actually needed? I assume this updates my location manager whenever the location changes, but it sure doesn't appear that there is much to the class itself! private class MyLocationListener implements LocationListener { @Override public void onLocationChanged(Location loc) { if (loc != null) { //Toast.makeText(getBaseContext(), "Location changed : Lat: " + loc.getLatitude() + " Lng: " + loc.getLongitude(), Toast.LENGTH_SHORT).show(); } } @Override public void onProviderDisabled(String provider) { // TODO Auto-generated method stub } @Override public void onProviderEnabled(String provider) { // TODO Auto-generated method stub } @Override public void onStatusChanged(String provider, int status, Bundle extras) { // TODO Auto-generated method stub } }

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  • Django Multi-Table Inheritance VS Specifying Explicit OneToOne Relationship in Models

    - by chefsmart
    Hope all this makes sense :) I'll clarify via comments if necessary. Also, I am experimenting using bold text in this question, and will edit it out if I (or you) find it distracting. With that out of the way... Using django.contrib.auth gives us User and Group, among other useful things that I can't do without (like basic messaging). In my app I have several different types of users. A user can be of only one type. That would easily be handled by groups, with a little extra care. However, these different users are related to each other in hierarchies / relationships. Let's take a look at these users: - Principals - "top level" users Administrators - each administrator reports to a Principal Coordinators - each coordinator reports to an Administrator Apart from these there are other user types that are not directly related, but may get related later on. For example, "Company" is another type of user, and can have various "Products", and products may be supervised by a "Coordinator". "Buyer" is another kind of user that may buy products. Now all these users have various other attributes, some of which are common to all types of users and some of which are distinct only to one user type. For example, all types of users have to have an address. On the other hand, only the Principal user belongs to a "BranchOffice". Another point, which was stated above, is that a User can only ever be of one type. The app also needs to keep track of who created and/or modified Principals, Administrators, Coordinators, Companies, Products etc. (So that's two more links to the User model.) In this scenario, is it a good idea to use Django's multi-table inheritance as follows: - from django.contrib.auth.models import User class Principal(User): # # # branchoffice = models.ForeignKey(BranchOffice) landline = models.CharField(blank=True, max_length=20) mobile = models.CharField(blank=True, max_length=20) created_by = models.ForeignKey(User, editable=False, blank=True, related_name="principalcreator") modified_by = models.ForeignKey(User, editable=False, blank=True, related_name="principalmodifier") # # # Or should I go about doing it like this: - class Principal(models.Model): # # # user = models.OneToOneField(User, blank=True) branchoffice = models.ForeignKey(BranchOffice) landline = models.CharField(blank=True, max_length=20) mobile = models.CharField(blank=True, max_length=20) created_by = models.ForeignKey(User, editable=False, blank=True, related_name="principalcreator") modified_by = models.ForeignKey(User, editable=False, blank=True, related_name="principalmodifier") # # # Please keep in mind that there are other user types that are related via foreign keys, for example: - class Administrator(models.Model): # # # principal = models.ForeignKey(Principal, help_text="The supervising principal for this Administrator") user = models.OneToOneField(User, blank=True) province = models.ForeignKey( Province) landline = models.CharField(blank=True, max_length=20) mobile = models.CharField(blank=True, max_length=20) created_by = models.ForeignKey(User, editable=False, blank=True, related_name="administratorcreator") modified_by = models.ForeignKey(User, editable=False, blank=True, related_name="administratormodifier") I am aware that Django does use a one-to-one relationship for multi-table inheritance behind the scenes. I am just not qualified enough to decide which is a more sound approach.

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  • How to combine designable components with dependency injection

    - by Wim Coenen
    When creating a designable .NET component, you are required to provide a default constructor. From the IComponent documentation: To be a component, a class must implement the IComponent interface and provide a basic constructor that requires no parameters or a single parameter of type IContainer. This makes it impossible to do dependency injection via constructor arguments. (Extra constructors could be provided, but the designer would ignore them.) Some alternatives we're considering: Service Locator Don't use dependency injection, instead use the service locator pattern to acquire dependencies. This seems to be what IComponent.Site.GetService is for. I guess we could create a reusable ISite implementation (ConfigurableServiceLocator?) which can be configured with the necessary dependencies. But how does this work in a designer context? Dependency Injection via properties Inject dependencies via properties. Provide default instances if they are necessary to show the component in a designer. Document which properties need to be injected. Inject dependencies with an Initialize method This is much like injection via properties but it keeps the list of dependencies that need to be injected in one place. This way the list of required dependencies is documented implicitly, and the compiler will assists you with errors when the list changes. Any idea what the best practice is here? How do you do it? edit: I have removed "(e.g. a WinForms UserControl)" since I intended the question to be about components in general. Components are all about inversion of control (see section 8.3.1 of the UMLv2 specification) so I don't think that "you shouldn't inject any services" is a good answer. edit 2: It took some playing with WPF and the MVVM pattern to finally "get" Mark's answer. I see now that visual controls are indeed a special case. As for using non-visual components on designer surfaces, I think the .NET component model is fundamentally incompatible with dependency injection. It appears to be designed around the service locator pattern instead. Maybe this will start to change with the infrastructure that was added in .NET 4.0 in the System.ComponentModel.Composition namespace.

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  • Passing Objects between different files

    - by user309779
    Typically, if I want to pass an object to an instance of something I would do it like so... Listing 1 File 1: public class SomeClass { // Some Properties public SomeClass() { public int ID { get { return mID; } set { mID = value; } } public string Name { set { mName = value; } get { return mName; } } } } public class SomeOtherClass { // Method 1 private void Method1(int one, int two) { SomeClass USER; // Create an instance Squid RsP = new Squid(); RsP.sqdReadUserConf(USER); // Able to pass 'USER' to class method in different file. } } In this example, I was not able to use the above approach. Probably because the above example passes an object between classes. Whereas, below, things are defined in a single class. I had to use some extra steps (trial & error) to get things to work. I am not sure what I did here or what its called. Is it good programming practice? Or is there is an easier way to do this (like above). Listing 2 File 1: private void SomeClass1 { [snip] TCOpt_fM.AutoUpdate = optAutoUpdate.Checked; TCOpt_fM.WhiteList = optWhiteList.Checked; TCOpt_fM.BlackList = optBlackList.Checked; [snip] private TCOpt TCOpt_fM; TCOpt_fM.SaveOptions(TCOpt_fM); } File 2: public class TCOpt: { public TCOpt OPTIONS; [snip] private bool mAutoUpdate = true; private bool mWhiteList = true; private bool mBlackList = true; [snip] public bool AutoUpdate { get { return mAutoUpdate; } set { mAutoUpdate = value; } } public bool WhiteList { get { return mWhiteList; } set { mWhiteList = value; } } public bool BlackList { get { return mBlackList; } set { mBlackList = value; } } [snip] public bool SaveOptions(TCOpt OPTIONS) { [snip] Some things being written out to a file here [snip] Squid soSwGP = new Squid(); soSgP.sqdWriteGlobalConf(OPTIONS); } } File 3: public class SomeClass2 { public bool sqdWriteGlobalConf(TCOpt OPTIONS) { Console.WriteLine(OPTIONS.WhiteSites); // Nothing prints here Console.WriteLine(OPTIONS.BlackSites); // Or here } } Thanks in advance, XO

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  • Can I call make runtime decided method calls in Java?

    - by Catalin Marin
    I know there is an invoke function that does the stuff, I am overall interested in the "correctness" of using such a behavior. My issue is this: I have a Service Object witch contains methods which I consider services. What I want to do is alter the behavior of those services without later intrusion. For example: class MyService { public ServiceResponse ServeMeDonuts() { do stuff... return new ServiceResponse(); } after 2 months I find out that I need to offer the same service to a new client app and I also need to do certain extra stuff like setting a flag, or make or updating certain data, or encode the response differently. What I can do is pop it up and throw down some IFs. In my opinion this is not good as it means interaction with tested code and may result in un wanted behaviour for the previous service clients. So I come and add something to my registry telling the system that the "NewClient" has a different behavior. So I'll do something like this: public interface Behavior { public void preExecute(); public void postExecute(); } public class BehaviorOfMyService implements Behavior{ String method; String clientType; public void BehaviorOfMyService(String method,String clientType) { this.method = method; this.clientType = clientType; } public void preExecute() { Method preCall = this.getClass().getMethod("pre" + this.method + this.clientType); if(preCall != null) { return preCall.invoke(); } return false; } ...same for postExecute(); public void preServeMeDonutsNewClient() { do the stuff... } } when the system will do something like this if(registrySaysThereIs different behavior set for this ServiceObject) { Class toBeCalled = Class.forName("BehaviorOf" + usedServiceObjectName); Object instance = toBeCalled.getConstructor().newInstance(method,client); instance.preExecute(); ....call the service... instance.postExecute(); .... } I am not particularly interested in correctness of code as in correctness of thinking and approach. Actually I have to do this in PHP, witch I see as a kind of Pop music of programming which I have to "sing" for commercial reasons, even though I play POP I really want to sing by the book, so putting aside my more or less inspired analogy I really want to know your opinion on this matter for it's practical necessity and technical approach. Thanks

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  • Autocomplete server-side implementation

    - by toluju
    What is a fast and efficient way to implement the server-side component for an autocomplete feature in an html input box? I am writing a service to autocomplete user queries in our web interface's main search box, and the completions are displayed in an ajax-powered dropdown. The data we are running queries against is simply a large table of concepts our system knows about, which matches roughly with the set of wikipedia page titles. For this service obviously speed is of utmost importance, as responsiveness of the web page is important to the user experience. The current implementation simply loads all concepts into memory in a sorted set, and performs a simple log(n) lookup on a user keystroke. The tailset is then used to provide additional matches beyond the closest match. The problem with this solution is that it does not scale. It currently is running up against the VM heap space limit (I've set -Xmx2g, which is about the most we can push on our 32 bit machines), and this prevents us from expanding our concept table or adding more functionality. Switching to 64-bit VMs on machines with more memory isn't an immediate option. I've been hesitant to start working on a disk-based solution as I am concerned that disk seek time will kill performance. Are there possible solutions that will let me scale better, either entirely in memory or with some fast disk-backed implementations? Edits: @Gandalf: For our use case it is important the the autocompletion is comprehensive and isn't just extra help for the user. As for what we are completing, it is a list of concept-type pairs. For example, possible entries are [("Microsoft", "Software Company"), ("Jeff Atwood", "Programmer"), ("StackOverflow.com", "Website")]. We are using Lucene for the full search once a user selects an item from the autocomplete list, but I am not yet sure Lucene would work well for the autocomplete itself. @Glen: No databases are being used here. When I'm talking about a table I just mean the structured representation of my data. @Jason Day: My original implementation to this problem was to use a Trie, but the memory bloat with that was actually worse than the sorted set due to needing a large number of object references. I'll read on the ternary search trees to see if it could be of use.

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  • Reorganizing MySQL table to multiple rows by timestamp.

    - by Ben Burleson
    OK MySQL Wizards: I have a table of position data from multiple probes defined as follows: +----------+----------+------+-----+---------+-------+ | Field | Type | Null | Key | Default | Extra | +----------+----------+------+-----+---------+-------+ | time | datetime | NO | | NULL | | | probe_id | char(3) | NO | | NULL | | | position | float | NO | | NULL | | +----------+----------+------+-----+---------+-------+ A simple select outputs something like this: +---------------------+----------+----------+ | time | probe_id | position | +---------------------+----------+----------+ | 2010-05-05 14:16:42 | 00A | 0.0045 | | 2010-05-05 14:16:42 | 00B | 0.0005 | | 2010-05-05 14:16:42 | 00C | 0.002 | | 2010-05-05 14:16:42 | 01A | 0 | | 2010-05-05 14:16:42 | 01B | 0.001 | | 2010-05-05 14:16:42 | 01C | 0.0025 | | 2010-05-05 14:16:43 | 00A | 0.0045 | | 2010-05-05 14:16:43 | 00B | 0.0005 | | 2010-05-05 14:16:43 | 00C | 0.002 | | 2010-05-05 14:16:43 | 01A | 0 | | . | . | . | | . | . | . | | . | . | . | +---------------------+----------+----------+ However, I'd like to output something like this: +---------------------+--------+--------+-------+-----+-------+--------+ | time | 00A | 00B | 00C | 01A | 01B | 01C | +---------------------+--------+--------+-------+-----+-------+--------+ | 2010-05-05 14:16:42 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | 2010-05-05 14:16:43 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | 2010-05-05 14:16:44 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | 2010-05-05 14:16:45 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | 2010-05-05 14:16:46 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | 2010-05-05 14:16:47 | 0.0045 | 0.0005 | 0.002 | 0 | 0.001 | 0.0025 | | . | . | . | . | . | . | . | | . | . | . | . | . | . | . | | . | . | . | . | . | . | . | +---------------------+--------+--------+-------+-----+-------+--------+ Ideally, the different probe position columns are dynamically generated based on data in the table. Is this possible, or am I pulling my hair out for nothing? I've tried GROUP BY time with GROUP_CONCAT that roughly gets the data out, but I can't separate that output into probe_id columns. mysql SELECT time, GROUP_CONCAT(probe_id), GROUP_CONCAT(position) FROM MG41 GROUP BY time LIMIT 10; +---------------------+-------------------------+------------------------------------+ | time | GROUP_CONCAT(probe_id) | GROUP_CONCAT(position) | +---------------------+-------------------------+------------------------------------+ | 2010-05-05 14:16:42 | 00A,00B,00C,01A,01B,01C | 0.0045,0.0005,0.002,0,0.001,0.0025 | | 2010-05-05 14:16:43 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:44 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:45 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:46 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:47 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:48 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:49 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:50 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | | 2010-05-05 14:16:51 | 01C,01B,01A,00C,00B,00A | 0.0025,0.001,0,0.002,0.0005,0.0045 | +---------------------+-------------------------+------------------------------------+

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  • Why does extend() engage in bizarre behaviour when passed the same list twice?

    - by intuited
    I'm pretty confused by one of the subtleties of the vimscript extend() function. If you use it to extend a list with another list, it does pretty much what you'd expect, which is to insert the second list into the first list at the index given by the third parameter: let list1 = [1,2,3,4,5,6] | echo extend(list1,[1,2,3,4,5,6],5) " [1, 2, 3, 4, 5, 1, 2, 3, 4, 5, 6, 6] However if you give it the same list twice it starts tripping out a bit. let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,0) " [1, 2, 3, 4, 5, 6, 1, 2, 3, 4, 5, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,1) " [1, 1, 1, 1, 1, 1, 1, 2, 3, 4, 5, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,2) " [1, 2, 1, 2, 1, 2, 1, 2, 3, 4, 5, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,3) " [1, 2, 3, 1, 2, 3, 1, 2, 3, 4, 5, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,4) " [1, 2, 3, 4, 1, 2, 3, 4, 1, 2, 5, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,5) " [1, 2, 3, 4, 5, 1, 2, 3, 4, 5, 1, 6] let list1 = [1,2,3,4,5,6] | echo extend(list1,list1,6) " [1, 2, 3, 4, 5, 6, 1, 2, 3, 4, 5, 6] Extra-confusingly, this behaviour applies when the list is referenced with two different variables: let list1 = [1,2,3,4,5,6] | let list2 = list1 | echo extend(list1,list2,4) " [1, 2, 3, 4, 1, 2, 3, 4, 1, 2, 5, 6] This is totally bizarre to me. I can't fathom a use for this functionality, and it seems like it would be really easy to invoke it by accident when you just wanted to insert one list into another and didn't realize that the variables were referencing the same array. The documentation says the following: If they are |Lists|: Append {expr2} to {expr1}. If {expr3} is given insert the items of {expr2} before item {expr3} in {expr1}. When {expr3} is zero insert before the first item. When {expr3} is equal to len({expr1}) then {expr2} is appended. Examples: :echo sort(extend(mylist, [7, 5])) :call extend(mylist, [2, 3], 1) When {expr1} is the same List as {expr2} then the number of items copied is equal to the original length of the List. E.g., when {expr3} is 1 you get N new copies of the first item (where N is the original length of the List). Does this make sense in a way that I'm not getting, or is it just an eccentricity?

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  • Choosing a distributed shared memory solution

    - by mindas
    I have a task to build a prototype for a massively scalable distributed shared memory (DSM) app. The prototype would only serve as a proof-of-concept, but I want to spend my time most effectively by picking the components which would be used in the real solution later on. The aim of this solution is to take data input from an external source, churn it and make the result available for a number of frontends. Those "frontends" would just take the data from the cache and serve it without extra processing. The amount of frontend hits on this data can literally be millions per second. The data itself is very volatile; it can (and does) change quite rapidly. However the frontends should see "old" data until the newest has been processed and cached. The processing and writing is done by a single (redundant) node while other nodes only read the data. In other words: no read-through behaviour. I was looking into solutions like memcached however this particular one doesn't fulfil all our requirements which are listed below: The solution must at least have Java client API which is reasonably well maintained as the rest of app is written in Java and we are seasoned Java developers; The solution must be totally elastic: it should be possible to add new nodes without restarting other nodes in the cluster; The solution must be able to handle failover. Yes, I realize this means some overhead, but the overall served data size isn't big (1G max) so this shouldn't be a problem. By "failover" I mean seamless execution without hardcoding/changing server IP address(es) like in memcached clients when a node goes down; Ideally it should be possible to specify the degree of data overlapping (e.g. how many copies of the same data should be stored in the DSM cluster); There is no need to permanently store all the data but there might be a need of post-processing of some of the data (e.g. serialization to the DB). Price. Obviously we prefer free/open source but we're happy to pay a reasonable amount if a solution is worth it. In any way, paid 24hr/day support contract is a must. The whole thing has to be hosted in our data centers so SaaS offerings like Amazon SimpleDB are out of scope. We would only consider this if no other options would be available. Ideally the solution would be strictly consistent (as in CAP); however, eventual consistence can be considered as an option. Thanks in advance for any ideas.

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  • Magento set Store Id - customer login - but still logged out

    - by user3564050
    I've got an overridden AccountController in which i set the current store to an other as currently running (example: Customer is in website default and store default, going to login page, click login, my loginPostAction sets the store to id "2" (on website 2) and then executes the parent code loginPostAction. The store is set, of course, but after the login and the redirect to home, the customer is not logged in anymore... Customer-sendlogindata-myaccountcontroller sets store-original account controller logs in without errors (cause $session customer is set)-redirect to home-customer is not logged in anymore... i set the store with Mage::app()-setCurrentStore($id); . And in index.php i've got an extra where the store is set to the right id (2) too and this works... but the customer is not logged in anymore.. is that an issue with the session cause different websites ? I don't want to globally share customer.. each website has his own customers, but every customer has to be able to login on default store. AccountController.php overridden: public $Website_Ids = array( array("code" => "gerstore", "id" => "3", "website" => "ger"), array("code" => "ukstore", "id" => "2", "website" => "uk"), array("code" => "esstore", "id" => "4", "website" => "es"), array("code" => "frstore", "id" => "5", "website" => "fr") ); public function loginPostAction() { $login = $this->getRequest()->get('login'); if(isset($login['username'])) { $found = null; foreach($this->Website_Ids as $WebsiteId) { $customer = Mage::getModel('customer/customer'); $customer->setWebsiteId($WebsiteId['id']); $customer->loadByEmail($login['username']); if(count($customer->getData()) > 0) { $found = $WebsiteId; } } if($found != null && Mage::app()->getStore()->getId() != $found['id']) { /* found, so set currentstore to id */ Mage::app()->setCurrentStore($found['id']); $_SESSION['current_store_b2b'] = $found; } /* not found, doesn't matter cause mage login exception handling */ } parent::loginPostAction(); } Index.php : session_start(); $session = $_SESSION['current_store_b2b']; if($session != null || $session != "") { Mage::app()->setCurrentStore($session['id']); Mage::run($session['code'], 'store'); } else { /* Store or website code */ $mageRunCode = isset($_SERVER['MAGE_RUN_CODE']) ? $_SERVER['MAGE_RUN_CODE'] : ''; /* Run store or run website */ $mageRunType = isset($_SERVER['MAGE_RUN_TYPE']) ? $_SERVER['MAGE_RUN_TYPE'] : 'store'; Mage::run($mageRunCode, $mageRunType); } Whats the matter ? Thanks.

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  • Who calls the Destructor of the class when operator delete is used in multiple inheritance.

    - by dicaprio-leonard
    This question may sound too silly, however , I don't find concrete answer any where else. With little knowledge on how late binding works and virtual keyword used in inheritance. As in the code sample, when in case of inheritance where a base class pointer pointing to a derived class object created on heap and delete operator is used to deallocate the memory , the destructor of the of the derived and base will be called in order only when the base destructor is declared virtual function. Now my question is : 1) When the destructor of base is not virtual, why the problem of not calling derived dtor occur only when in case of using "delete" operator , why not in the case given below: derived drvd; base *bPtr; bPtr = &drvd; //DTOR called in proper order when goes out of scope. 2) When "delete" operator is used, who is reponsible to call the destructor of the class? The operator delete will have an implementation to call the DTOR ? or complier writes some extra stuff ? If the operator has the implementation then how does it looks like , [I need sample code how this would have been implemented]. 3) If virtual keyword is used in this example, how does operator delete now know which DTOR to call? Fundamentaly i want to know who calls the dtor of the class when delete is used. Sample Code class base { public: base() { cout<<"Base CTOR called"<<endl; } virtual ~base() { cout<<"Base DTOR called"<<endl; } }; class derived:public base { public: derived() { cout<<"Derived CTOR called"<<endl; } ~derived() { cout<<"Derived DTOR called"<<endl; } }; I'm not sure if this is a duplicate, I couldn't find in search. int main() { base *bPtr = new derived(); delete bPtr;// only when you explicitly try to delete an object return 0; }

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  • User created Validator wont call Client side validation Javascript on 'complex' user control.

    Hi All, I have created a user control (from System.Web.UI.UserControl), and created my own validator for the user control (from System.Web.UI.WebControls.BaseValidator). Everything works ok until I try to get the user control to do client side validation. While trying to debug this issue I have set 'Control to Validate' to a text box instead of the custom user control, and the client side script works fine! It appears to me that it has an a issue with my composite user control I have created. Has anyone encountered this issue before? Has anyone else seen client side validation fail on custom user controls? Some extra info : The composite control is a drop down list and 'loader image', as it is a ajax enabled drop down list (using ICallbackEventHandler). I know that the client side javascript is being written to the page, and have placed an alert('random message') as the first line in the validator function that only appears if it is validating a text box (i.e. not when it is validating my custom control) Language : C# (ASP.NET 2.0) and jQuery 1.2.6 in aspx file : <rms:UserDDL ID="ddlUserTypes" runat="server" PreLoad="true" /> <rms:DDLValidator ID="userTypesVal" ControlToValidate="ddlUserTypes" ErrorMessage="You have not selected a UserType" runat="server" Text="You have not selected a UserType" Display="Dynamic" EnableClientScript="true" /> in validator code behind protected string ScriptBlock { get { string nl = System.Environment.NewLine; return "<script type=\"text/javascript\">" + nl + " function " + ScriptBlockFunctionName + "(ctrl)" + nl + " {" + nl + " alert('Random message'); " + nl + " var selVal = $('#' + ctrl.controltovalidate).val(); " + nl + " alert(selVal);" + nl + " if (selVal === '-1') return false; " + nl + " return false; " + nl + " }" + nl + "</script>"; } } protected override void OnPreRender(EventArgs e) { if (this.DetermineRenderUplevel() && this.EnableClientScript) { Page.ClientScript.RegisterExpandoAttribute(this.ClientID, "evaluationfunction", this.ScriptBlockFunctionName); Page.ClientScript.RegisterClientScriptBlock(GetType(), this.ScriptBlockKey, this.ScriptBlock); } base.OnPreRender(e); } I know my ControlPropertiesValid() and EvaluateIsValid() work ok. I appreciate any help on this issue. Noel.

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  • How to put a background image on GridBagLayout

    - by Loligans
    I am trying to work with layout managers for the first time, and they are just kicking me in the teeth. I am trying to make a background image and then put buttons on top, using GridBagLayout, if there is a a better layoutmanager please do tell. As for trying to learn how to use layout managers, its very difficult and any learning references would also be much appreciated. This is what it looks like currently, I can get the frame to show the full image, but when i use gridlayout manager, it does that public void addComponentsToPane(Container pane){ BackgroundImage image = new BackgroundImage(); JButton button1, button2, button3, button4, button5; pane.setLayout(new GridBagLayout()); GridBagConstraints c = new GridBagConstraints(); if(shouldFill){ c.fill = GridBagConstraints.NONE; } button1 = new JButton("Button 1"); if (shouldWeightX) { c.weightx = 0.5; } c.fill = GridBagConstraints.HORIZONTAL; c.gridx = 1; c.gridy = 0; button1.setOpaque(false); pane.add(button1, c); button2 = new JButton("Button 2"); c.fill = GridBagConstraints.HORIZONTAL; c.weightx = 0.5; c.gridx = 0; c.gridy = 0; button2.setOpaque(false); pane.add(button2, c); button3 = new JButton("Button 3"); c.fill = GridBagConstraints.HORIZONTAL; c.weightx = 0.5; c.gridx = 2; c.gridy = 0; button3.setOpaque(false); pane.add(button3, c); button4 = new JButton("Long-Named Button 4"); c.fill = GridBagConstraints.HORIZONTAL; c.ipady = 40; //make this component tall c.weightx = 0.0; c.gridwidth = 3; c.gridx = 0; c.gridy = 1; pane.add(button4, c); button5 = new JButton("button 1"); c.fill = GridBagConstraints.HORIZONTAL; c.ipady = 0; //reset to default c.weighty = 1.0; //request any extra vertical space c.anchor = GridBagConstraints.PAGE_END; //bottom of space c.insets = new Insets(10,0,0,0); //top padding c.gridx = 1; //aligned with button 2 c.gridwidth = 2; //2 columns wide c.gridy = 2; //third row pane.add(button5, c); c.ipadx = 800; c.ipady = 400; pane.add(image, c); } This is what i'm trying to make it look like

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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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  • Overwrite values when using gridview Edit?

    - by sah302
    I am using a GridView which is bound to a LinqDataSource and using the automatic edit and delete buttons. However, I don't want the user to edit two of the columns manually, but done automatically. Specifically username who last updated the entry, and the date it was updated. The gridview only contains 3 columns: Name, Date modified, last updated by. Right now when the user clicks the edit button they can only edit the name column (other two set to read-only). Upon clicking the update button, I want the other 2 fields to update as well based on some extra code. I thought this was done in the code behind within the event rowUpdating, but it doesn't seem to work. My gridview: <asp:GridView ID="gvNewsSources" runat="server" AutoGenerateColumns="False" DataSourceID="ldsNewsSource" AutoGenerateDeleteButton="True" AutoGenerateEditButton="True" CellPadding="4" ForeColor="#333333" GridLines="None" DataKeyNames="Id"> <RowStyle BackColor="#F7F6F3" ForeColor="#333333" /> <Columns> <asp:BoundField DataField="Name" HeaderText="Name" SortExpression="Name" /> <asp:BoundField DataField="LastUpdatedBy" HeaderText="Last Updated By" SortExpression="LastUpdatedBy" ReadOnly="True" /> <asp:BoundField DataField="DatedModified" HeaderText="Dated Modified" SortExpression="DatedModified" ReadOnly="True" /> </Columns> <FooterStyle BackColor="#5D7B9D" Font-Bold="True" ForeColor="White" /> <PagerStyle BackColor="#284775" ForeColor="White" HorizontalAlign="Center" /> <SelectedRowStyle BackColor="#E2DED6" Font-Bold="True" ForeColor="#333333" /> <HeaderStyle BackColor="#5D7B9D" Font-Bold="True" ForeColor="White" /> <EditRowStyle BackColor="#999999" /> <AlternatingRowStyle BackColor="White" ForeColor="#284775" /> </asp:GridView> My code behind: Partial Class _Default Inherits System.Web.UI.Page Protected Sub gvNewsSources_RowUpdating(ByVal sender As Object, ByVal e As System.Web.UI.WebControls.GridViewUpdateEventArgs) Handles gvNewsSources.RowUpdating e.NewValues("LastUpdatedBy") = GetUser.GetUserName e.NewValues("DateModified") = Date.Now() lblOutput.Text = e.NewValues("DateModified").ToString() End Sub End Class Yet when I run through this, I get no errors, but the values aren't being updated in the database or in the gridview. I ran through debug mode and the new values dictionary starts at 1 and ends up being 3 by the end of the rowUpdating event and the value is being set (tested by output the newValue of Datemodified), but it isn't saving. What am I doing wrong?

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  • Is it ok to dynamic cast "this" as a return value?

    - by Panayiotis Karabassis
    This is more of a design question. I have a template class, and I want to add extra methods to it depending on the template type. To practice the DRY principle, I have come up with this pattern (definitions intentionally omitted): template <class T> class BaseVector: public boost::array<T, 3> { protected: BaseVector<T>(const T x, const T y, const T z); public: bool operator == (const Vector<T> &other) const; Vector<T> operator + (const Vector<T> &other) const; Vector<T> operator - (const Vector<T> &other) const; Vector<T> &operator += (const Vector<T> &other) { (*this)[0] += other[0]; (*this)[1] += other[1]; (*this)[2] += other[2]; return *dynamic_cast<Vector<T> * const>(this); } } template <class T> class Vector : public BaseVector<T> { public: Vector<T>(const T x, const T y, const T z) : BaseVector<T>(x, y, z) { } }; template <> class Vector<double> : public BaseVector<double> { public: Vector<double>(const double x, const double y, const double z); Vector<double>(const Vector<int> &other); double norm() const; }; I intend BaseVector to be nothing more than an implementation detail. This works, but I am concerned about operator+=. My question is: is the dynamic cast of the this pointer a code smell? Is there a better way to achieve what I am trying to do (avoid code duplication, and unnecessary casts in the user code)? Or am I safe since, the BaseVector constructor is private?

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