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  • Simple Project Templates

    - by Geertjan
    The NetBeans sources include a module named "simple.project.templates": In the module sources, Tim Boudreau turns out to be the author of the code, so I asked him what it was all about, and if he could provide some usage code. His response, from approximately this time last year because it's been sitting in my inbox for a while, is below. Sure - though I think the javadoc in it is fairly complete.  I wrote it because I needed to create a bunch of project templates for Javacard, and all of the ways that is usually done were grotesque and complicated.  I figured we already have the ability to create files from templates, and we already have the ability to do substitutions in templates, so why not have a single file that defines the project as a list of file templates to create (with substitutions in the names) and some definitions of what should be in project properties. You can also add files to the project programmatically if you want.Basically, a template for an entire project is a .properties file.  Any line which doesn't have the prefix 'pp.' or 'pvp.' is treated as the definition of one file which should be created in the new project.  Any such line where the key ends in * means that file should be opened once the new project is created.  So, for example, in the nodejs module, the definition looks like: {{projectName}}.js*=Templates/javascript/HelloWorld.js .npmignore=node_hidden_templates/npmignore So, the first line means:  - Create a file with the same name as the project, using the HelloWorld template    - I.e. the left side of the line is the relative path of the file to create, and the right side is the path in the system filesystem for the template to use       - If the template is not one you normally want users to see, just register it in the system filesystem somewhere other than Templates/ (but remember to set the attribute that marks it as a template)  - Include that file in the set of files which should be opened in the editor once the new project is created. To actually create a project, first you just create a new ProjectCreator: ProjectCreator gen = new ProjectCreator( parentFolderOfNewProject ); Now, if you want to programmatically generate any files, in addition to those defined in the template, you can: gen.add (new FileCreator("nbproject", "project.xml", false) {     public DataObject create (FileObject project, Map<String,String> substitutions) throws IOException {          ...     } }); Then pass the FileObject for the project template (the properties file) to the ProjectCreator's createProject method (hmm, maybe it should be the string path to the project template instead, to save the caller trouble looking up the FileObject for the template).  That method looks like this: public final GeneratedProject createProject(final ProgressHandle handle, final String name, final FileObject template, final Map<String, String> substitutions) throws IOException { The name parameter should be the directory name for the new project;  the map is the strings you gathered in the wizard which should be used for substitutions.  createProject should be called on a background thread (i.e. use a ProgressInstantiatingIterator for the wizard iterator and just pass in the ProgressHandle you are given). The return value is a GeneratedProject object, which is just a holder for the created project directory and the set of DataObjects which should be opened when the wizard finishes. I'd love to see simple.project.templates moved out of the javacard cluster, as it is really useful and much simpler than any of the stuff currently done for generating projects.  It would also be possible to do much richer tools for creating projects in apisupport - i.e. choose (or create in the wizard) the templates you want to use, generate a skeleton wizard with a UI for all the properties you'd like to substitute, etc. Here is a partial project template from Javacard - for example usage, see org.netbeans.modules.javacard.wizard.ProjectWizardIterator in javacard.project (or the much simpler one in contrib/nodejs). #This properties file describes what to create when a project template is#instantiated.  The keys are paths on disk relative to the project root. #The values are paths to the templates to use for those files in the system#filesystem.  Any string inside {{ and }}'s will be substituted using properties#gathered in the template wizard.#Special key prefixes are #  pp. - indicates an entry for nbproject/project.properties#  pvp. - indicates an entry for nbproject/private/private.properties #File templates, in format [path-in-project=path-to-template]META-INF/javacard.xml=org-netbeans-modules-javacard/templates/javacard.xmlMETA-INF/MANIFEST.MF=org-netbeans-modules-javacard/templates/EAP_MANIFEST.MF APPLET-INF/applet.xml=org-netbeans-modules-javacard/templates/applet.xmlscripts/{{classnamelowercase}}.scr=org-netbeans-modules-javacard/templates/test.scrsrc/{{packagepath}}/{{classname}}.java*=Templates/javacard/ExtendedApplet.java nbproject/deployment.xml=org-netbeans-modules-javacard/templates/deployment.xml#project.properties contentspp.display.name={{projectname}}pp.platform.active={{activeplatform}} pp.active.device={{activedevice}}pp.includes=**pp.excludes= I will be using the above info in an upcoming blog entry and provide step by step instructions showing how to use them. However, anyone else out there should have enough info from the above to get started yourself!

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  • Subterranean IL: Generics and array covariance

    - by Simon Cooper
    Arrays in .NET are curious beasts. They are the only built-in collection types in the CLR, and SZ-arrays (single dimension, zero-indexed) have their own commands and IL syntax. One of their stranger properties is they have a kind of built-in covariance long before generic variance was added in .NET 4. However, this causes a subtle but important problem with generics. First of all, we need to briefly recap on array covariance. SZ-array covariance To demonstrate, I'll tweak the classes I introduced in my previous posts: public class IncrementableClass { public int Value; public virtual void Increment(int incrementBy) { Value += incrementBy; } } public class IncrementableClassx2 : IncrementableClass { public override void Increment(int incrementBy) { base.Increment(incrementBy); base.Increment(incrementBy); } } In the CLR, SZ-arrays of reference types are implicitly convertible to arrays of the element's supertypes, all the way up to object (note that this does not apply to value types). That is, an instance of IncrementableClassx2[] can be used wherever a IncrementableClass[] or object[] is required. When an SZ-array could be used in this fashion, a run-time type check is performed when you try to insert an object into the array to make sure you're not trying to insert an instance of IncrementableClass into an IncrementableClassx2[]. This check means that the following code will compile fine but will fail at run-time: IncrementableClass[] array = new IncrementableClassx2[1]; array[0] = new IncrementableClass(); // throws ArrayTypeMismatchException These checks are enforced by the various stelem* and ldelem* il instructions in such a way as to ensure you can't insert a IncrementableClass into a IncrementableClassx2[]. For the rest of this post, however, I'm going to concentrate on the ldelema instruction. ldelema This instruction pops the array index (int32) and array reference (O) off the stack, and pushes a pointer (&) to the corresponding array element. However, unlike the ldelem instruction, the instruction's type argument must match the run-time array type exactly. This is because, once you've got a managed pointer, you can use that pointer to both load and store values in that array element using the ldind* and stind* (load/store indirect) instructions. As the same pointer can be used for both input and output to the array, the type argument to ldelema must be invariant. At the time, this was a perfectly reasonable restriction, and maintained array type-safety within managed code. However, along came generics, and with it the constrained callvirt instruction. So, what happens when we combine array covariance and constrained callvirt? .method public static void CallIncrementArrayValue() { // IncrementableClassx2[] arr = new IncrementableClassx2[1] ldc.i4.1 newarr IncrementableClassx2 // arr[0] = new IncrementableClassx2(); dup newobj instance void IncrementableClassx2::.ctor() ldc.i4.0 stelem.ref // IncrementArrayValue<IncrementableClass>(arr, 0) // here, we're treating an IncrementableClassx2[] as IncrementableClass[] dup ldc.i4.0 call void IncrementArrayValue<class IncrementableClass>(!!0[],int32) // ... ret } .method public static void IncrementArrayValue<(IncrementableClass) T>( !!T[] arr, int32 index) { // arr[index].Increment(1) ldarg.0 ldarg.1 ldelema !!T ldc.i4.1 constrained. !!T callvirt instance void IIncrementable::Increment(int32) ret } And the result: Unhandled Exception: System.ArrayTypeMismatchException: Attempted to access an element as a type incompatible with the array. at IncrementArrayValue[T](T[] arr, Int32 index) at CallIncrementArrayValue() Hmm. We're instantiating the generic method as IncrementArrayValue<IncrementableClass>, but passing in an IncrementableClassx2[], hence the ldelema instruction is failing as it's expecting an IncrementableClass[]. On features and feature conflicts What we've got here is a conflict between existing behaviour (ldelema ensuring type safety on covariant arrays) and new behaviour (managed pointers to object references used for every constrained callvirt on generic type instances). And, although this is an edge case, there is no general workaround. The generic method could be hidden behind several layers of assemblies, wrappers and interfaces that make it a requirement to use array covariance when calling the generic method. Furthermore, this will only fail at runtime, whereas compile-time safety is what generics were designed for! The solution is the readonly. prefix instruction. This modifies the ldelema instruction to ignore the exact type check for arrays of reference types, and so it lets us take the address of array elements using a covariant type to the actual run-time type of the array: .method public static void IncrementArrayValue<(IncrementableClass) T>( !!T[] arr, int32 index) { // arr[index].Increment(1) ldarg.0 ldarg.1 readonly. ldelema !!T ldc.i4.1 constrained. !!T callvirt instance void IIncrementable::Increment(int32) ret } But what about type safety? In return for ignoring the type check, the resulting controlled mutability pointer can only be used in the following situations: As the object parameter to ldfld, ldflda, stfld, call and constrained callvirt instructions As the pointer parameter to ldobj or ldind* As the source parameter to cpobj In other words, the only operations allowed are those that read from the pointer; stind* and similar that alter the pointer itself are banned. This ensures that the array element we're pointing to won't be changed to anything untoward, and so type safety within the array is maintained. This is a typical example of the maxim that whenever you add a feature to a program, you have to consider how that feature interacts with every single one of the existing features. Although an edge case, the readonly. prefix instruction ensures that generics and array covariance work together and that compile-time type safety is maintained. Tune in next time for a look at the .ctor generic type constraint, and what it means.

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  • University teaches DOS-style C++, how to deal with it

    - by gaidal
    Half a year ago I had a look at available programming educations. I chose this one because unlike most of the choices: The majority of the courses seemed to be about something concrete and useful; the languages used are C++ and Java which are platform-independent; later courses include developing for mobile devices and a course on Android development, which seemed modern and relevant. Now after two introductory courses we're just starting with C++, and my programming professor seems a bit weird. He's tested us on things like "why should you use constants" and "why are globals bad" in a kind of mechanical way, without much context, before teaching actual programming. His handouts use system("pause"), system("cls"), and getch() from some conio.h that seems ancient according to what I've read. I just did a task that was about printing the "ASCII letters from 32 to 255" (huh?), with an example picture showing a table with Windows' Extended ASCII - of course I got other results for 128-255 on my Arch Linux that uses Unicode, and this isn't mentioned at all. I don't know, it just doesn't seem right... As if he is teaching programming because he has to, perhaps? Should I bring such things up? Hmm. I was looking forward to learning from someone who really knows stuff, and in an academic, rigorous way, like SICP or something. Aren't professors in programming supposed to be like that? I studied math for a while and every teacher and assistant there were really precise about what they said, but this is my second programming teacher that is sort of disappointing. Oh well. Now, question: Is this what to expect from universities or Not OK, and how do I deal with it? I have never touched the language C++ (or C) until now, and am not the right person to jump up and say "This is So Wrong!", so if I google something and find 10 people who say "xxx is blasphemy", how do I skillfully communicate this? I do think it would be better for those classmates who are total beginners not to learn bad habits (such as these vibes of total ignorance of other platforms!) during the upcoming courses, but don't want to disrespect the teacher. I don't know if it's reasonable or just cocky to bring up things like "what about other platforms?" or "but what about this article or stackoverflow answer that I read that said..." for every assignment? Or, if he keeps ignoring non-Windows-programming, should I give up and focus on my own projects or somehow argue that this really isn't OK nowadays? Are there any programming teachers out there, what do you think? By the way these are web-based courses, all interaction between teachers and students takes place in a forum. EDIT: A few answers seem to be making some incorrect assumptions, so maybe I should add a few things. I have been doing programming for fun on and off for 10 years, am pretty comfortable in 3 languages and read programming blogs et c regularly. Also, I feel kind of done being a student, having a degree in another field. I just need another, relevant diploma to work as a programmer, so I'm going back for that. Studying computer science for 5 years is not for me anymore, even though I enjoy learning and solving problems in my free time. Second, let me highlight that I don't expect it to be like the industry at all, quite the contrary. I expect it to be academic, dry and unnecessarily correct. No, it's not just math. Every professor I have had in math, or Japanese (major) or Chinese (minor) have been very very academic, discussing subtle points for hours with passion. But the courses I'm taking now and a previous one in programming don't seem serious. They neither resemble industry NOR academia. That is the problem. And it's not because I can't learn programming anyway. Third, I don't necessarily want to learn C++ or Android development, and I know I could teach myself those and anything else if I wanted to. But I am going back to school anyway, and those platform-independent languages and mobile stuff made me think that maybe they're serious about teaching something relevant here. Seems like I got this wrong, but we'll see.

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  • What Counts for A DBA - Logic

    - by drsql
    "There are 10 kinds of people in the world. Those who will always wonder why there are only two items in my list and those who will figured it out the first time they saw this very old joke."  Those readers who will give up immediately and get frustrated with me for not explaining it to them are not likely going to be great technical professionals of any sort, much less a programmer or administrator who will be constantly dealing with the common failures that make up a DBA's day.  Many of these people will stare at this like a dog staring at a traffic signal and still have no more idea of how to decipher the riddle. Without explanation they will give up, call the joke "stupid" and, feeling quite superior, walk away indignantly to their job likely flipping patties of meat-by-product. As a data professional or any programmer who has strayed  to this very data-oriented blog, you would, if you are worth your weight in air, either have recognized immediately what was going on, or felt a bit ignorant.  Your friends are chuckling over the joke, but why is it funny? Unfortunately you left your smartphone at home on the dresser because you were up late last night programming and were running late to work (again), so you will either have to fake a laugh or figure it out.  Digging through the joke, you figure out that the word "two" is the most important part, since initially the joke mentioned 10. Hmm, why did they spell out two, but not ten? Maybe 10 could be interpreted a different way?  As a DBA, this sort of logic comes into play every day, and sometimes it doesn't involve nerdy riddles or Star Wars folklore.  When you turn on your computer and get the dreaded blue screen of death, you don't immediately cry to the help desk and sit on your thumbs and whine about not being able to work. Do that and your co-workers will question your nerd-hood; I know I certainly would. You figure out the problem, and when you have it narrowed down, you call the help desk and tell them what the problem is, usually having to explain that yes, you did in fact try to reboot before calling.  Of course, sometimes humility does come in to play when you reach the end of your abilities, but the ‘end of abilities’ is not something any of us recognize readily. It is handy to have the ability to use logic to solve uncommon problems: It becomes especially useful when you are trying to solve a data-related problem such as a query performance issue, and the way that you approach things will tell your coworkers a great deal about your abilities.  The novice is likely to immediately take the approach of  trying to add more indexes or blaming the hardware. As you become more and more experienced, it becomes increasingly obvious that performance issues are a very complex topic. A query may be slow for a myriad of reasons, from concurrency issues, a poor query plan because of a parameter value (like parameter sniffing,) poor coding standards, or just because it is a complex query that is going to be slow sometimes. Some queries that you will deal with may have twenty joins and hundreds of search criteria, and it can take a lot of thought to determine what is going on.  You can usually figure out the problem to almost any query by using basic knowledge of how joins and queries work, together with the help of such things as the query plan, profiler or monitoring tools.  It is not unlikely that it can take a full day’s work to understand some queries, breaking them down into smaller queries to find a very tiny problem. Not every time will you actually find the problem, and it is part of the process to occasionally admit that the problem is random, and everything works fine now.  Sometimes, it is necessary to realize that a problem is outside of your current knowledge, and admit temporary defeat: You can, at least, narrow down the source of the problem by looking logically at all of the possible solutions. By doing this, you can satisfy your curiosity and learn more about what the actual problem was. For example, in the joke, had you never been exposed to the concept of binary numbers, there is no way you could have known that binary - 10 = decimal - 2, but you could have logically come to the conclusion that 10 must not mean ten in the context of the joke, and at that point you are that much closer to getting the joke and at least won't feel so ignorant.

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  • Android Marketplace Error: "The server could not process your apk. Try again."

    - by jdandrea
    I have an updated apk - tested successfully on various devices and simulator instances - with the following manifest: <?xml version="1.0" encoding="utf-8"?> <manifest xmlns:android="http://schemas.android.com/apk/res/android" package="com.myCompany.appName" android:versionCode="2" android:versionName="1.0.1"> <uses-sdk android:minSdkVersion="3" android:targetSdkVersion="5" /> <uses-permission android:name="android.permission.INTERNET" /> <supports-screens android:largeScreens="true" android:normalScreens="true" android:smallScreens="true" /> <application android:icon="@drawable/icon" android:label="@string/icon_name" android:debuggable="false"> <activity android:name=".myActivity" android:configChanges="keyboardHidden|orientation"> <intent-filter> <action android:name="android.intent.action.MAIN" /> <category android:name="android.intent.category.LAUNCHER" /> </intent-filter> </activity> </application> </manifest> When I post to Android Marketplace as an upgrade to my existing 1.0 app, I get the aforementioned ambiguous message: "The server could not process your apk. Try again." I've searched elsewhere for this message in hopes of finding out what might be happening, to no avail. (A popular suggestion is to move the uses-sdk element to the top of the manifest, but as you can see it's already at the top.) Clues welcome/appreciated. Update: I just tried to upload the same file again. Now I get a new message: The new apk's versionCode (2) in AndroidManifest.xml must be higher than the old apk's versionCode (2). The server could not process your apk. Try again. Soooo Marketplace did get my upgraded apk after all? (The very first accepted apk's versionCode was 1, so this update was of course bumped to 2.) Confused … Bumping it up to 3 and trying again. Surprise surprise, I get the original "could not process" error all over again. Going in circles. Hmm ... :( Nuther Update: If I exit and re-enter the Marketplace page, now it shows that the app has been uploaded! Except there's no app icon. Curiouser and curiouser ... and this is all happening with a cache-cleared (standards-friendly) browser to boot. So - do I trust the upload? Or start over ... with versionCode="4"? All I want is to get a solid "Upload successful, here's the icon, ready to publish" type of response.

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  • Problems with makeObjectsPerformSelector inside and outside a class?

    - by QuakAttak
    A friend and I are creating a card game for the iPhone, and in these early days of the project, I'm developing a Deck class and a Card class to keep up with the cards. I'm wanting to test the shuffle method of the Deck class, but I am not able to show the values of the cards in the Deck class instance. The Deck class has a NSArray of Card objects that have a method called displayCard that shows the value and suit using console output(printf or NSLog). In order to show what cards are in a Deck instance all at once, I am using this, [deck makeObjectsPerformSelector:@selector(displayCard)], where deck is the NSArray in the Deck class. Inside of the Deck class, nothing is displayed on the console output. But in a test file, it works just fine. Here's the test file that creates its own NSArray: #import <Foundation/Foundation.h> #import "card.h" int main (int argc, char** argv) { NSAutoreleasePool* pool = [[NSAutoreleasePool alloc] init]; Card* two = [[Card alloc] initValue:2 withSuit:'d']; Card* three = [[Card alloc] initValue:3 withSuit:'h']; Card* four = [[Card alloc] initValue:4 withSuit:'c']; NSArray* deck = [NSArray arrayWithObjects:two,three,four,nil]; //Ok, what if we release the objects in the array before they're used? //I don't think this will work... [two release]; [three release]; [four release]; //Ok, it works... I wonder how... //Hmm... how will this work? [deck makeObjectsPerformSelector:@selector(displayCard)]; //Yay! It works fine! [pool release]; return 0; } This worked beautifully, so I created an initializer around this idea, creating 52 card objects one at a time and adding them to the NSArray using deck = [deck arrayByAddingObject:newCard]. Is the real problem with how I'm using makeObjectsPerformSelector or something before/after it?

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  • LINQ to SQL: To Attach or Not To Attach

    - by bradhe
    So I'm have a really hard time figuring out when I should be attaching to an object and when I shouldn't be attaching to an object. First thing's first, here is a small diagram of my (very simplified) object model. Edit: Okay, apparently I'm not allowed to post images...here you go: http://i.imgur.com/2ROFI.png In my DAL I create a new DataContext every time I do a data-related operation. Say, for instance, I want to save a new user. In my business layer I create a new user. var user = new User(); user.FirstName = "Bob"; user.LastName = "Smith"; user.Username = "bob.smith"; user.Password = StringUtilities.EncodePassword("MyPassword123"); user.Organization = someOrganization; // Assume that someOrganization was loaded and it's data context has been garbage collected. Now I want to go save this user. var userRepository = new RepositoryFactory.GetRepository<UserRepository>(); userRepository.Save(user); Neato! Here is my save logic: public void Save(User user) { if (!DataContext.Users.Contains(user)) { user.Id = Guid.NewGuid(); user.CreatedDate = DateTime.Now; user.Disabled = false; //DataContext.Organizations.Attach(user.Organization); DataContext.Users.InsertOnSubmit(user); } else { DataContext.Users.Attach(user); } DataContext.SubmitChanges(); // Finished here as well. user.Detach(); } So, here we are. You'll notice that I comment out the bit where the DataContext attachs to the organization. If I attach to the organization I get the following exception: NotSupportedException: An attempt has been made to Attach or Add an entity that is not new, perhaps having been loaded from another DataContext. This is not supported. Hmm, that doesn't work. Let me try it without attaching (i.e. comment out that line about attaching to the organization). DuplicateKeyException: Cannot add an entity with a key that is already in use. WHAAAAT? I can only assume this is trying to insert a new organization which is obviously false. So, what's the deal guys? What should I do? What is the proper approach? It seems like L2S makes this quite a bit harder than it should be...

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  • Use a subdirectory as root with htaccess in Apache 1.3

    - by Andrew
    I'm trying to deploy a site generated with Jekyll and would like to keep the site in its own subfolder on my server to keep everything more organized. Essentially, I'd like to use the contents of /jekyll as the root unless a file similarly named exists in the actual web root. So something like /jekyll/sample-page/ would show as http://www.example.com/sample-page/, while something like /other-folder/ would display as http://www.example.com/other-folder. My test server runs Apache 2.2 and the following .htaccess (adapted from http://gist.github.com/97822) works flawlessly: RewriteEngine On # Map http://www.example.com to /jekyll. RewriteRule ^$ /jekyll/ [L] # Map http://www.example.com/x to /jekyll/x unless there is a x in the web root. RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteCond %{REQUEST_URI} !^/jekyll/ RewriteRule ^(.*)$ /jekyll/$1 # Add trailing slash to directories without them so DirectoryIndex works. # This does not expose the internal URL. RewriteCond %{REQUEST_FILENAME} -d RewriteCond %{REQUEST_FILENAME} !/$ RewriteRule ^(.*)$ $1/ # Disable auto-adding slashes to directories without them, since this happens # after mod_rewrite and exposes the rewritten internal URL, e.g. turning # http://www.example.com/about into http://www.example.com/jekyll/about. DirectorySlash off However, my production server runs Apache 1.3, which doesn't allow DirectorySlash. If I disable it, the server gives a 500 error because of internal redirect overload. If I comment out the last section of ReWriteConds and rules: RewriteCond %{REQUEST_FILENAME} -d RewriteCond %{REQUEST_FILENAME} !/$ RewriteRule ^(.*)$ $1/ …everything mostly works: http://www.example.com/sample-page/ displays the correct content. However, if I omit the trailing slash, the URL in the address bar exposes the real internal URL structure: http://www.example.com/jekyll/sample-page/ What is the best way to account for directory slashes in Apache 1.3, where useful tools like DirectorySlash don't exist? How can I use the /jekyll/ directory as the site root without revealing the actual URL structure? Edit: After a ton of research into Apache 1.3, I've found that this problem is essentially a combination of two different issues listed at the Apache 1.3 URL Rewriting Guide. I have a (partially) moved DocumentRoot, which in theory would be taken care of with something like this: RewriteRule ^/$ /e/www/ [R] I also have the infamous "Trailing Slash Problem," which is solved by setting the RewriteBase (as was suggested in one of the responses below): RewriteBase /~quux/ RewriteRule ^foo$ foo/ [R] The problem is combining the two. Moving the document root doesn't (can't?) use RewriteBase—fixing trailing slashes requires(?) it… Hmm…

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  • Problem getting ar_mailer/ar_sendmail working on new server

    - by Max Williams
    Hey all. I've got a new app up and running on a new ubuntu server. It's working fine generally but i can't get ar_sendmail working. I'm following the instructions on this page: http://www.ameravant.com/posts/sending-tons-of-emails-in-ruby-on-rails-with-ar_mailer The setup is all done, ie i can "deliver mails" which just saves records in my Email table. Now i want to get the ar_sendmail daemon running to actually send them. (so i'm at 'Running ar_sendmail in daemon mode' in that web page). First thing: ar_sendmail --mailq >>ar_sendmail: command not found Ok...so, where is ar_sendmail? I have a look and there's an ar_sendmail file in the bin folder of the ar_mailer plugin, so i add the location of that to my path. I don't know if this was the right thing to do or not. Ok, so try again. ar_sendmail --mailq /var/www/apps/millionaire/vendor/plugins/ar_mailer/bin/ar_sendmail:3:in `require': no such file to load -- action_mailer/ar_sendmail (LoadError) from /var/www/apps/millionaire/vendor/plugins/ar_mailer/bin/ar_sendmail:3 hmm. Here's the offending file, there's not much there. #!/usr/bin/env ruby require 'action_mailer/ar_sendmail' ActionMailer::ARSendmail.run ok...so it literally is just trying to require this and can't find it. The file, action_mailer/ar_sendmail.rb is in the ar_mailer plugin, in it's lib folder. So, given that it's being called from inside the plugin, it should be able to see this right? I've got a feeling that i'm way off the track here and have missed something simple. Can anyone set me straight? I'm using rails 2.3.4 in case that's relevant. EDIT - i just realised something kind of dumb: when i call ar_sendmail from the command line like this, i'm just loading that one file, which doesn't know where it's supposed to look for the rest of the stuff, i think. Which really makes me think that i'm not trying to run the right thing. Is the ar_sendmail daemon a seperate program altogether, that i would get with apt_get or something? EDIT2 - i made some progress by installing the ar_mailer gem (which the guide said i shouldn't do) and that does seem to run. It's sending some mail request somewhere and clearing the Email table of pending emails. Running ar_sendmail in -ov (oneshot verbal) mode i see it report this for example: sent email 00000000019 from [email protected] to [email protected]: # So, it actually looks like it's working now and i just need to set up the ACTUAL THING WHICH SENDS EMAILS. sigh. still grateful for any advice. thanks, max

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  • NSDocument Subclass not closed by NSWindowController?

    - by Nathan Douglas
    Okay, I'm fairly new to Cocoa and Objective-C, and to OOP in general. As background, I'm working on an extensible editor that stores the user's documents in a package. This of course required some "fun" to get around some issues with NSFileWrapper (i.e. a somewhat sneaky writing and loading process to avoid making NSFileWrappers for every single document within the bundle). The solution I arrived at was to essentially treat my NSDocument subclass as just a shell -- use it to make the folder for the bundle, and then pass off writing the actual content of the document to other methods. Unfortunately, at some point I seem to have completely screwed the pooch. I don't know how this happened, but closing the document window no longer releases the document. The document object doesn't seem to receive a "close" message -- or any related messages -- even though the window closes successfully. The end result is that if I start my app, create a new document, save it, then close it, and try to reopen it, the document window never appears. With some creative subclassing and NSLogging, I managed to figure out that the document object was still in memory, and still attached to the NSDocumentController instance, and so trying to open the document never got past the NSDocumentController's "hmm, currently have that one open" check. I did have an NSWindowController and NSDocumentController instance, but I've purged them from my project completely. I've overridden nearly every method for NSDocument trying to find out where the issue is. So far as I know, my Interface Builder bindings are all correct -- "Close" in the main menu is attached to "performClose:" of the First Responder, etc, and I've tried with fresh unsullied MainMenu and Document xibs as well. I thought that it might be something strange with my bundle writing code, so I basically deleted it all and started from scratch, but that didn't seem to work. I took out my init method overrides, and that didn't help either. I don't have the source of any simple document apps here, so I didn't try the next logical step (to substitute known-working code for mine in the readfromurl and writetourl methods). I've had this problem for about sixteen hours of uninterrupted troubleshooting now, and needless to say, I'm at the end of my rope. If I can't figure it out, I guess I'm going to try the project from scratch with a lot more code and intensity based around the bundle-document mess. Any help would be greatly appreciated.

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  • Coldfusion Report Builder - How can you set different datasources externally between prod/staging/de

    - by Smooth Operator
    Coldfusion Report Builder is great. One small issue. We use ANT+CFANT to deploy. When we create the report, say in a datasource called MyApp_dev on a dev box. Everything works great when the report is created. We deploy the report to our staging server, which has a datasource of MyApp_Staging. That server also, may or may not, have the live app working under MyApp_Live. Ant pushes the update to Staging just great. Run the report, crashes and burns. Why? It seems the report is looking for the MyApp_Dev data_source, even though the application is using the MyApp_Staging datasource. In digging around I found a few approaches, I would like to do this one, final, ideal way from the beginning instead of having to go back to do dozens of reports differently when I have a new Aha! moment. 1) Obvious: Pass in the datasource in to the cfreport tag. Doesn't work for ColdFusion Builder Reports as of v8, or v9 as tested on Linux. 2) Most realistic option (but painful) so far: Pass in the query as an object into the ColdFusion Builder report. Let's think about this: Create the Report with the report builder to my heart's content using the RDS, etc on my local box. When I'm done, copy the query into a snippet of code, or into a database column to be dynamically be injected at runtime with correct datasource. Modify my "run report" event to find the query from the database column, insert it into another dynamic cfquery and potentially... evaluate (!?!) it? Fun side is I can set the cfquery datasource to what I would need for each environment. When I modify the report's columns in CF Report Builder, I always have to update the query in the database. Is there a snippet of code that can extract this for me? Hmm. 3) Less than ideal. Suck it up and let all the reports in staging run off the live server. Maybe copy the live data into staging (sans structural changes) to let it seem similar. Are there any eloquent ways to accomplish the above? Thanks in Advance!

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  • What to name 2 methods with same signatures

    - by coffeeaddict
    Initially I had a method in our DL that would take in the object it's updating like so: internal void UpdateCash(Cash Cash) { using (OurCustomDbConnection conn = CreateConnection("UpdateCash")) { conn.CommandText = @"update Cash set captureID = @captureID, ac_code = @acCode, captureDate = @captureDate, errmsg = @errorMessage, isDebit = @isDebit, SourceInfoID = @sourceInfoID, PayPalTransactionInfoID = @payPalTransactionInfoID, CreditCardTransactionInfoID = @CreditCardTransactionInfoID where id = @cashID"; conn.AddParam("@captureID", cash.CaptureID); conn.AddParam("@acCode", cash.ActionCode); conn.AddParam("@captureDate", cash.CaptureDate); conn.AddParam("@errorMessage", cash.ErrorMessage); conn.AddParam("@isDebit", cyberCash.IsDebit); conn.AddParam("@PayPalTransactionInfoID", cash.PayPalTransactionInfoID); conn.AddParam("@CreditCardTransactionInfoID", cash.CreditCardTransactionInfoID); conn.AddParam("@sourceInfoID", cash.SourceInfoID); conn.AddParam("@cashID", cash.Id); conn.ExecuteNonQuery(); } } My boss felt that creating an object every time just to update one or two fields is overkill. But I had a couple places in code using this. He recommended using just UpdateCash and sending in the ID for CAsh and field I want to update. Well the problem is I have 2 places in code using my original method. And those 2 places are updating 2 completely different fields in the Cash table. Before I was just able to get the existing Cash record and shove it into a Cash object, then update the properties I wanted to be updated in the DB, then send back the cash object to my method above. I need some advice on what to do here. I have 2 methods and they have the same signature. I'm not quite sure what to rename these because both are updating 2 completely different fields in the Cash table: internal void UpdateCash(int cashID, int paypalCaptureID) { using (OurCustomDbConnection conn = CreateConnection("UpdateCash")) { conn.CommandText = @"update Cash set CaptureID = @paypalCaptureID where id = @cashID"; conn.AddParam("@captureID", paypalCaptureID); conn.ExecuteNonQuery(); } } internal void UpdateCash(int cashID, int PayPalTransactionInfoID) { using (OurCustomDbConnection conn = CreateConnection("UpdateCash")) { conn.CommandText = @"update Cash set PaymentSourceID = @PayPalTransactionInfoID where id = @cashID"; conn.AddParam("@PayPalTransactionInfoID", PayPalTransactionInfoID); conn.ExecuteNonQuery(); } } So I thought hmm, maybe change the names to these so that they are now unique and somewhat explain what field its updating: UpdateCashOrderID UpdateCashTransactionInfoID ok but that's not really very good names. And I can't go too generic, for example: UpdateCashTransaction(int cashID, paypalTransactionID) What if we have different types of transactionIDs that the cash record holds besides just the paypalTransactionInfoID? such as the creditCardInfoID? Then what? Transaction doesn't tell me what kind. And furthermore what if you're updating 2 fields so you have 2 params next to the cashID param: UpdateCashTransaction(int cashID, paypalTransactionID, someOtherFieldIWantToUpdate) see my frustration? what's the best way to handle this is my boss doesn't like my first route?

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  • Accessing contents of NativeWindow in a HTML AIR application?

    - by Dan Scotton
    I'm currently building a HTML/JS AIR application. The application needs to display to the user a different 'window' - dependant on whether this is the first time they've launched the application or not. This part is actually fine and I have the code below to do that: if(!startUp()) { // this simply returns a boolean from a local preferences database that gets shipped with the .air // do first time stuff var windowOptions = new air.NativeWindowInitOptions(); windowOptions.systemChrome = 'none'; windowOptions.type = 'lightweight'; windowOptions.transparent = 'true'; windowOptions.resizable = 'false'; var windowBounds = new air.Rectangle(300, 300, 596, 490); var newHtmlLoader = air.HTMLLoader.createRootWindow(true, windowOptions, true, windowBounds); newHtmlLoader.load(new air.URLRequest('cover.html')); } else { // display default window // just set nativeWindow.visible = true (loaded from application.xml) } However, what I want to be able to do is manipulate the html content from within cover.html after it has loaded up. There seems to be plenty of tutorials online of how to move, resize, etc. the NativeWindow, but I simply want access to the NativeWindow's HTML content. For example, how would I add a new paragraph to that page? I've tried the following: newHtmlLoader.window.opener = window; var doc = newHtmlLoader.window.opener.document.documentElement; Using AIR's Introspector console, ....log(doc) returns [object HTMLHtmlElement]. Hmm, seems promising right? I then go on to try: var p = document.createElement('p'); var t = document.createTextNode('Insert Me'); p.appendChild(t); doc.appendChild(p); ...but nothing gets inserted. I've also tried the following replacements for doc: var doc = newHtmlLoader.window.opener.document.body; // .log(doc) -> [object HTMLBodyElement] var doc = newHtmlLoader.window.opener.document; // .log(doc) -> Error: HIERARCHY_REQUEST_ERR: DOM Exception 3 ...as well as the following with jQuery: $(doc).append('<p>Insert Me</p>'); // again, nothing So, anyone had any experience in accessing a NativeWindow's inner content programmatically? Any help will be greatly appreciated.

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  • Writing a managed wrapper for unmanaged (C++) code - custom types/structs

    - by Bobby
    faacEncConfigurationPtr FAACAPI faacEncGetCurrentConfiguration( faacEncHandle hEncoder); I'm trying to come up with a simple wrapper for this C++ library; I've never done more than very simple p/invoke interop before - like one function call with primitive arguments. So, given the above C++ function, for example, what should I do to deal with the return type, and parameter? FAACAPI is defined as: #define FAACAPI __stdcall faacEncConfigurationPtr is defined: typedef struct faacEncConfiguration { int version; char *name; char *copyright; unsigned int mpegVersion; unsigned long bitRate; unsigned int inputFormat; int shortctl; psymodellist_t *psymodellist; int channel_map[64]; } faacEncConfiguration, *faacEncConfigurationPtr; AFAIK this means that the return type of the function is a reference to this struct? And faacEncHandle is: typedef struct { unsigned int numChannels; unsigned long sampleRate; ... SR_INFO *srInfo; double *sampleBuff[MAX_CHANNELS]; ... double *freqBuff[MAX_CHANNELS]; double *overlapBuff[MAX_CHANNELS]; double *msSpectrum[MAX_CHANNELS]; CoderInfo coderInfo[MAX_CHANNELS]; ChannelInfo channelInfo[MAX_CHANNELS]; PsyInfo psyInfo[MAX_CHANNELS]; GlobalPsyInfo gpsyInfo; faacEncConfiguration config; psymodel_t *psymodel; /* quantizer specific config */ AACQuantCfg aacquantCfg; /* FFT Tables */ FFT_Tables fft_tables; int bitDiff; } faacEncStruct, *faacEncHandle; So within that struct we see a lot of other types... hmm. Essentially, I'm trying to figure out how to deal with these types in my managed wrapper? Do I need to create versions of these types/structs, in C#? Something like this: [StructLayout(LayoutKind.Sequential)] struct faacEncConfiguration { uint useTns; ulong bitRate; ... } If so then can the runtime automatically "map" these objects onto eachother? And, would I have to create these "mapped" types for all the types in these return types/parameter type hierarchies, all the way down until I get to all primitives? I know this is a broad topic, any advice on getting up-to-speed quickly on what I need to learn to make this happen would be very much appreciated! Thanks!

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  • jQuery .load(), don't show new content until images loaded

    - by Jarred
    Hi. I have been working on a jQuery photo slideshow. It scales the images to the browser size, and slides them left and right. There is no pre-determined size or aspect ratio, the script does everything on the fly. It requires that all images be fully loaded, so it can custom resize each individual image based on it's own aspect ratio ( width():height(), etc ), calculate the width of containing div, and calculate the slide distance from one image to another. As a stand-alone, it works pretty well (despite my lack of skills)! I simply hide the slideshow containing div at (document).ready, allow the images to load, then run the slideshow prep scripts at (window).load. Once this is done, it only then makes the slideshow divs, images, etc appear, properly sized, positioned and ready to roll. The ultimate goal is to be able to load in any number of slideshows without refreshing the page. The point of this is to be able to play uninterrupted background music. I know music on websites is annoying, but the target market likes it, a lot! I am using (target).load(page.php .element, function prepInsertNewShow() { //Prepare slideshow resizeImages(); slideArray(); //Show slideshow (target).fadeIn(); }); and it definitely works! The problem is that I cannot find a way to hold off on preparing and showing the new content until the images have finished loading. It is running the slideshow prep scripts (which are totally dependent on the images being fully loaded), before the images are loaded. This results in a completely jacked up show! What I want to do is this - (target).load(page.php .element, function prepInsertNewShow() { //Wait until images are loaded $('img').load( function() { //Prepare slideshow resizeImages(); slideArray(); //Show slideshow (target).fadeIn(); } }); But this doesn't seem to work, the new content is never shown. You can see a live version here. The initial gallery loads correctly, everything looks good. The only nav link that works is Galleries Engagement, which will load a new show (a containing div with multiple <img> tags). You will see that the images are not centered, the containing div and slide distances are much too small, as they were calculated using images that were not actually loaded. Is there any way I can delay handling and showing new content until it is fully loaded? Any suggestions would be most appreciated, thanks for your time! PS - It just occurred to me while typing this that a decent solution may be to insert "width=x" height="x" into the <img> tags, so the script can work from those values, even if the images have not loaded...hmm...

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  • I cut-to-move DCIM folder to ext SD when an auto android OS update popped up b4 I could choose target - Cannot recover 200+ photos

    - by ZeroG
    I was downloading my Exhibit II's DCIM camera folder (with month's of photos inside) to its external SD card, in order to transfer them into my laptop. In my overconfidence, I hurriedly chose cut-to-move (rather than copy-to-move) when KABOOM! —an automatic Android OS update popped up before I could choose the target!!! I figured everything was in cache & calmly tried to go through with the update. But that was not a typically seamless event. It showed downloading icon but hmm… since I rooted the phone it brought the command line up & recovery sequence. But neither Android nor I had yet downloaded any alternate custom ROM Files to internal SD to update from! So were they trying to make me unroot my phone by giving me some bogus update on the fly or just give me a hard time in trying to hand me down an unrooted ROM that I'd have to figure out how to root again? Yes, I know there was that blurb about overwriting a file of the same name but I was trying to shake the darn stubborn update being forced on my phone during this precarious moment. I thought I had frozen or turned off all those auto-updates previously. Anyway, phones are small & fingers are big (sigh)... I tried to reboot into safe mode but the resultant photo file was partially overwritten (200 files had names but Zero bytes in them). I thought maybe it was still hung in cache or deposited somewhere else but I have searched everywhere with file managers. Since I did not have Titanium backing up camera, photo folder or gallery, I cannot recover 200+ photos. Dumb. You can understand my dilemma as I am involved in the arts & although just a camera phone, most of these photos were historic & aesthetic or at least as to subject matter. Photo-ops don't reoccur. I have tried a couple of recovery apps from the market like Search Duplicates & Recover to no avail. I was only able to salvage stuff I'd sent out in messages. I've got several decades in computers & this is such a miserable beginner's piece of bad luck I can't believe it happened to me. They were precious photos! Yes, I turned on Titanium since & yes I even tried USB to laptop recoveries. Being on a MacBookPro I'm trying androidfiletransfer.dmg, but I'd have to upgrade to Peach Sunrise to get above Android 3.0 for that App to recognize the phone via USB & the programmer says installation zeros your data, so that pretty much toasts any secret hidden places where these photos may have been deposited. Don't want to do that & am still trying to find them. They certainly didn't make it to my external SD Card. If any of you techies out there know anything, please help & thanks. Despite decades of being in computing, unfamiliar & ever-changing hard or software can humble even the most seasoned veterans.

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  • Gaming blew fuse and causes funny smell: how to overcome?

    - by George Tomlinson
    I've been gaming for a while now. When playing certain games this PC goes into overdrive. The fan/fans start/s to sound like a jet engine it/they get/s so busy. Also I have smelt burning when this has happened. The fuse blew on the 4 socket adapter I was using recently. On the following thread someone said this could be due to the PSU not being strong enough to handle the load, in what it seems could be a related issue someone had, although the person who posted this question did say that blowing a fan on their PC stopped it crashing in that case: http://www.tomshardware.co.uk/answers/id-2047543/gtx-650-overheating-issue.html. This is exactly what they said: Your GPU isn't overheating. 70+ before it would shutdown and cause a restart. Make sure your PSU is strong enough to handle your new system at load and possibly run Memtest to check your RAM (although not BSOD'ing and just shutting down points to the PSU). This (the PSU part) makes more sense to me than it being to do with dust etc, since it seems a more plausible explanation of why the fuse blew. The PC has no problems except when playing certain games: i.e. TERA Rising and WoW with add-ons (I think WoW is ok as long as I don't have more than 1 add-on (Healers Have To Die)). I'm just wondering if anyone knows or can suggest what I might be able to do to be able to play these games without this problem occurring. The PC's spec is this: Display: NVIDIA GeForce GTX 650 8GB RAM (6 available) Processor: AMD FX (tm) - 8120 Eight-Core Processor - 3.1 GHz, 4 Cores, 8 Logical Processors I have read on another post that forcing vsync in the Nvidia Control Panel helped with what seems could be a similar problem, so I plan to see if that solves it, God permitting. EDIT: I tried the Vsync thing, and it seems the situation may have improved, although this may be due to something else: i.e. maybe the PC was working harder yesterday, due to just having downloaded a few things or lots of things running. I'm still noticing the funny smell when playing TERA. It's not so much burning: it's more like glue. The smell might have had a burning element to it in the past, but I think it's always had a glue element. EDIT 2: the PSU is an 'ATX Switching Power Supply', Model E-500ATX. Other info it gives on the PSU is 230V, Current 10A and Frequency 50-60Hz. It also has some other info which I can supply if necessary. Putting the PC plug in the wall socket instead of the power strip seems like it might have reduced the load on the PC quite a bit: I think it sounds less stressed. it has been off for a while whilst I took the side panel off though, so I'll wait to see what happens before getting too excited. EDIT 3: hmm. So here's the latest: just playing TERA. The fan's running quite fast again. Hard to tell whether switching to the wall socket has made a difference in terms of strain on the PC: I don't know if one would expect it to. Still seems like it might have helped though. Oh and there didn't seem to be much dust in the PC, although I didn't disconnect any components. I'm still getting the glue type smell. ASIDE: reminds me of someone on a PC near me at the library once who was actually sniffing glue right there in front of everyone while on the PC and he started talking about how he was sniffing glue. lol. That's no joke. EDIT 4: So the questions now are: Question 1: Is the smell something I should sort out? (If so, how might I do this?) Question 2: is it necessary to take any steps to prevent blowing another fuse (and if so which step/s?).

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  • Security Trimmed Cross Site Collection Navigation

    - by Sahil Malik
    Ad:: SharePoint 2007 Training in .NET 3.5 technologies (more information). This article will serve as documentation of a fully functional codeplex project that I just created. This project will give you a WebPart that will give you security trimmed navigation across site collections. The first question is, why create such a project? In every single SharePoint project you will do, one question you will always be faced with is, what should the boundaries of sites be, and what should the boundaries of site collections be? There is no good or bad answer to this, because it really really depends on your needs. There are some factors in play here. Site Collections will allow you to scale, as a Site collection is the smallest entity you can put inside a content database Site collections will allow you to offer different levels of SLAs, because you put a site collection on a separate content database, and put that database on a separate server. Site collections are a security boundary – and they can be moved around at will without affecting other site collections. Site collections are also a branding boundary. They are also a feature deployment boundary, so you can have two site collections on the same web application with completely different nature of services. But site collections break navigation, i.e. a site collection at “/”, and a site collection at “/sites/mySiteCollection”, are completely independent of each other. If you have access to both, the navigation of / won’t show you a link to /sites/mySiteCollection. Some people refer to this as a huge issue in SharePoint. Luckily, some workarounds exist. A long time ago, I had blogged about “Implementing Consistent Navigation across Site Collections”. That approach was a no-code solution, it worked – it gave you a consistent navigation across site collections. But, it didn’t work in a security trimmed fashion! i.e., if I don’t have access to Site Collection ‘X’, it would still show me a link to ‘X’. Well this project gets around that issue. Simply deploy this project, and it’ll give you a WebPart. You can use that WebPart as either a webpart or as a server control dropped via SharePoint designer, and it will give you Security Trimmed Cross Site Collection Navigation. The code has been written for SP2010, but it will work in SP2007 with the help of http://spwcfsupport.codeplex.com . What do I need to do to make it work? I’m glad you asked! Simple! Deploy the .wsp (which you can download here). This will give you a site collection feature called “Winsmarts Cross Site Collection Navigation” as shown below. Go ahead and activate it, and this will give you a WebPart called “Winsmarts Navigation Web Part” as shown below: Just drop this WebPart on your page, and it will show you all site collections that the currently logged in user has access to. Really it’s that easy! This is shown as below - In the above example, I have two site collections that I created at /sites/SiteCollection1 and /sites/SiteCollection2. The navigation shows the titles. You see some extraneous crap as well, you might want to clean that – I’ll talk about that in a minute. What? You’re running into problems? If the problem you’re running into is that you are prompted to login three times, and then it shows a blank webpart that says “Loading your applications ..” and then craps out!, then most probably you’re using a different authentication scheme. Behind the scenes I use a custom WCF service to perform this job. OOTB, I’ve set it to work with NTLM, but if you need to make it work alternate authentications such as forms based auth, or client side certs, you will need to edit the %14%\ISAPI\Winsmarts.CrossSCNav\web.config file, specifically, this section - 1: <bindings> 2: <webHttpBinding> 3: <binding name="customWebHttpBinding"> 4: <security mode="TransportCredentialOnly"> 5: <transport clientCredentialType="Ntlm"/> 6: </security> 7: </binding> 8: </webHttpBinding> 9: </bindings> For Kerberos, change the “clientCredentialType” to “Windows” For Forms auth, remove that transport line For client certs – well that’s a bit more involved, but it’s just web.config changes – hit a good book on WCF or hire me for a billion trillion $. But fair warning, I might be too busy to help immediately. If you’re running into a different problem, please leave a comment below, but the code is pretty rock solid, so .. hmm .. check what you’re doing! BTW, I don’t  make any guarantee/warranty on this – if this code makes you sterile, unpopular, bad hairstyle, anything else, that is your problem! But, there are some known issues - I wrote this as a concept – you can easily extend it to be more flexible. Example, hierarchical nav, or, horizontal nav, jazzy effects with jquery or silverlight– all those are possible very very easily. This webpart is not smart enough to co-exist with another instance of itself on the same page. I can easily extend it to do so, which I will do in my spare(!?) time! Okay good! But that’s not all! As you can see, just dropping the WebPart may show you many extraneous site collections, or maybe you want to restrict which site collections are shown, or exclude a certain site collection to be shown from the navigation. To support that, I created a property on the WebPart called “UrlMatchPattern”, which is a regex expression you specify to trim the results :). So, just edit the WebPart, and specify a string property of “http://sp2010/sites/” as shown below. Note that you can put in whatever regex expression you want! So go crazy, I don’t care! And this gives you a cleaner look.   w00t! Enjoy! Comment on the article ....

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  • Analytic functions – they’re not aggregates

    - by Rob Farley
    SQL 2012 brings us a bunch of new analytic functions, together with enhancements to the OVER clause. People who have known me over the years will remember that I’m a big fan of the OVER clause and the types of things that it brings us when applied to aggregate functions, as well as the ranking functions that it enables. The OVER clause was introduced in SQL Server 2005, and remained frustratingly unchanged until SQL Server 2012. This post is going to look at a particular aspect of the analytic functions though (not the enhancements to the OVER clause). When I give presentations about the analytic functions around Australia as part of the tour of SQL Saturdays (starting in Brisbane this Thursday), and in Chicago next month, I’ll make sure it’s sufficiently well described. But for this post – I’m going to skip that and assume you get it. The analytic functions introduced in SQL 2012 seem to come in pairs – FIRST_VALUE and LAST_VALUE, LAG and LEAD, CUME_DIST and PERCENT_RANK, PERCENTILE_CONT and PERCENTILE_DISC. Perhaps frustratingly, they take slightly different forms as well. The ones I want to look at now are FIRST_VALUE and LAST_VALUE, and PERCENTILE_CONT and PERCENTILE_DISC. The reason I’m pulling this ones out is that they always produce the same result within their partitions (if you’re applying them to the whole partition). Consider the following query: SELECT     YEAR(OrderDate),     FIRST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING),     LAST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING),     PERCENTILE_CONT(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)),     PERCENTILE_DISC(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)) FROM Sales.SalesOrderHeader ; This is designed to get the TotalDue for the first order of the year, the last order of the year, and also the 95% percentile, using both the continuous and discrete methods (‘discrete’ means it picks the closest one from the values available – ‘continuous’ means it will happily use something between, similar to what you would do for a traditional median of four values). I’m sure you can imagine the results – a different value for each field, but within each year, all the rows the same. Notice that I’m not grouping by the year. Nor am I filtering. This query gives us a result for every row in the SalesOrderHeader table – 31465 in this case (using the original AdventureWorks that dates back to the SQL 2005 days). The RANGE BETWEEN bit in FIRST_VALUE and LAST_VALUE is needed to make sure that we’re considering all the rows available. If we don’t specify that, it assumes we only mean “RANGE BETWEEN UNBOUNDED PRECEDING AND CURRENT ROW”, which means that LAST_VALUE ends up being the row we’re looking at. At this point you might think about other environments such as Access or Reporting Services, and remember aggregate functions like FIRST. We really should be able to do something like: SELECT     YEAR(OrderDate),     FIRST_VALUE(TotalDue)         OVER (PARTITION BY YEAR(OrderDate)               ORDER BY OrderDate, SalesOrderID               RANGE BETWEEN UNBOUNDED PRECEDING                         AND UNBOUNDED FOLLOWING) FROM Sales.SalesOrderHeader GROUP BY YEAR(OrderDate) ; But you can’t. You get that age-old error: Msg 8120, Level 16, State 1, Line 5 Column 'Sales.SalesOrderHeader.OrderDate' is invalid in the select list because it is not contained in either an aggregate function or the GROUP BY clause. Msg 8120, Level 16, State 1, Line 5 Column 'Sales.SalesOrderHeader.SalesOrderID' is invalid in the select list because it is not contained in either an aggregate function or the GROUP BY clause. Hmm. You see, FIRST_VALUE isn’t an aggregate function. None of these analytic functions are. There are too many things involved for SQL to realise that the values produced might be identical within the group. Furthermore, you can’t even surround it in a MAX. Then you get a different error, telling you that you can’t use windowed functions in the context of an aggregate. And so we end up grouping by doing a DISTINCT. SELECT DISTINCT     YEAR(OrderDate),         FIRST_VALUE(TotalDue)              OVER (PARTITION BY YEAR(OrderDate)                   ORDER BY OrderDate, SalesOrderID                   RANGE BETWEEN UNBOUNDED PRECEDING                             AND UNBOUNDED FOLLOWING),         LAST_VALUE(TotalDue)             OVER (PARTITION BY YEAR(OrderDate)                   ORDER BY OrderDate, SalesOrderID                   RANGE BETWEEN UNBOUNDED PRECEDING                             AND UNBOUNDED FOLLOWING),     PERCENTILE_CONT(0.95)          WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)),     PERCENTILE_DISC(0.95)         WITHIN GROUP (ORDER BY TotalDue)         OVER (PARTITION BY YEAR(OrderDate)) FROM Sales.SalesOrderHeader ; I’m sorry. It’s just the way it goes. Hopefully it’ll change the future, but for now, it’s what you’ll have to do. If we look in the execution plan, we see that it’s incredibly ugly, and actually works out the results of these analytic functions for all 31465 rows, finally performing the distinct operation to convert it into the four rows we get in the results. You might be able to achieve a better plan using things like TOP, or the kind of calculation that I used in http://sqlblog.com/blogs/rob_farley/archive/2011/08/23/t-sql-thoughts-about-the-95th-percentile.aspx (which is how PERCENTILE_CONT works), but it’s definitely convenient to use these functions, and in time, I’m sure we’ll see good improvements in the way that they are implemented. Oh, and this post should be good for fellow SQL Server MVP Nigel Sammy’s T-SQL Tuesday this month.

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  • Networking in VirtualBox

    - by Fat Bloke
    Networking in VirtualBox is extremely powerful, but can also be a bit daunting, so here's a quick overview of the different ways you can setup networking in VirtualBox, with a few pointers as to which configurations should be used and when. VirtualBox allows you to configure up to 8 virtual NICs (Network Interface Controllers) for each guest vm (although only 4 are exposed in the GUI) and for each of these NICs you can configure: Which virtualized NIC-type is exposed to the Guest. Examples include: Intel PRO/1000 MT Server (82545EM),  AMD PCNet FAST III (Am79C973, the default) or  a Paravirtualized network adapter (virtio-net). How the NIC operates with respect to your Host's physical networking. The main modes are: Network Address Translation (NAT) Bridged networking Internal networking Host-only networking NAT with Port-forwarding The choice of NIC-type comes down to whether the guest has drivers for that NIC.  VirtualBox, suggests a NIC based on the guest OS-type that you specify during creation of the vm, and you rarely need to modify this. But the choice of networking mode depends on how you want to use your vm (client or server) and whether you want other machines on your network to see it. So let's look at each mode in a bit more detail... Network Address Translation (NAT) This is the default mode for new vm's and works great in most situations when the Guest is a "client" type of vm. (i.e. most network connections are outbound). Here's how it works: When the guest OS boots,  it typically uses DHCP to get an IP address. VirtualBox will field this DHCP request and tell the guest OS its assigned IP address and the gateway address for routing outbound connections. In this mode, every vm is assigned the same IP address (10.0.2.15) because each vm thinks they are on their own isolated network. And when they send their traffic via the gateway (10.0.2.2) VirtualBox rewrites the packets to make them appear as though they originated from the Host, rather than the Guest (running inside the Host). This means that the Guest will work even as the Host moves from network to network (e.g. laptop moving between locations), and from wireless to wired connections too. However, how does another computer initiate a connection into a Guest?  e.g. connecting to a web server running in the Guest. This is not (normally) possible using NAT mode as there is no route into the Guest OS. So for vm's running servers we need a different networking mode.... Bridged Networking Bridged Networking is used when you want your vm to be a full network citizen, i.e. to be an equal to your host machine on the network. In this mode, a virtual NIC is "bridged" to a physical NIC on your host, like this: The effect of this is that each VM has access to the physical network in the same way as your host. It can access any service on the network such as external DHCP services, name lookup services, and routing information just as the host does. Logically, the network looks like this: The downside of this mode is that if you run many vm's you can quickly run out of IP addresses or your network administrator gets fed up with you asking for statically assigned IP addresses. Secondly, if your host has multiple physical NICs (e.g. Wireless and Wired) you must reconfigure the bridge when your host jumps networks.  Hmm, so what if you want to run servers in vm's but don't want to involve your network administrator? Maybe one of the next 2 modes is for you... Internal Networking When you configure one or more vm's to sit on an Internal network, VirtualBox ensures that all traffic on that network stays within the host and is only visible to vm's on that virtual network. Configuration looks like this: The internal network ( in this example "intnet" ) is a totally isolated network and so is very "quiet". This is good for testing when you need a separate, clean network, and you can create sophisticated internal networks with vm's that provide their own services to the internal network. (e.g. Active Directory, DHCP, etc). Note that not even the Host is a member of the internal network, but this mode allows vm's to function even when the Host is not connected to a network (e.g. on a plane). Note that in this mode, VirtualBox provides no "convenience" services such as DHCP, so your machines must be statically configured or one of the vm's needs to provide a DHCP/Name service. Multiple internal networks are possible and you can configure vm's to have multiple NICs to sit across internal and other network modes and thereby provide routes if needed. But all this sounds tricky. What if you want an Internal Network that the host participates on with VirtualBox providing IP addresses to the Guests? Ah, then for this, you might want to consider Host-only Networking... Host-only Networking Host-only Networking is like Internal Networking in that you indicate which network the Guest sits on, in this case, "vboxnet0": All vm's sitting on this "vboxnet0" network will see each other, and additionally, the host can see these vm's too. However, other external machines cannot see Guests on this network, hence the name "Host-only". Logically, the network looks like this: This looks very similar to Internal Networking but the host is now on "vboxnet0" and can provide DHCP services. To configure how a Host-only network behaves, look in the VirtualBox Manager...Preferences...Network dialog: Port-Forwarding with NAT Networking Now you may think that we've provided enough modes here to handle every eventuality but here's just one more... What if you cart around a mobile-demo or dev environment on, say, a laptop and you have one or more vm's that you need other machines to connect into? And you are continually hopping onto different (customer?) networks. In this scenario: NAT - won't work because external machines need to connect in. Bridged - possibly an option, but does your customer want you eating IP addresses and can your software cope with changing networks? Internal - we need the vm(s) to be visible on the network, so this is no good. Host-only - same problem as above, we want external machines to connect in to the vm's. Enter Port-forwarding to save the day! Configure your vm's to use NAT networking; Add Port Forwarding rules; External machines connect to "host":"port number" and connections are forwarded by VirtualBox to the guest:port number specified. For example, if your vm runs a web server on port 80, you could set up rules like this:  ...which reads: "any connections on port 8080 on the Host will be forwarded onto this vm's port 80".  This provides a mobile demo system which won't need re-configuring every time you open your laptop lid. Summary VirtualBox has a very powerful set of options allowing you to set up almost any configuration your heart desires. For more information, check out the VirtualBox User Manual on Virtual Networking. -FB 

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  • XBRL US Conference Highlights

    - by john.orourke(at)oracle.com
    Back in early November I had an opportunity to attend the XBRL US National Conference in Philadelphia.  At the event, XBRL US announced that Oracle had joined the initiative, so I had a chance to participate in a press conference and attend a number of sessions.  Oracle joined XBRL US so we can stay ahead of the standard and leverage it in our products, and to help drive awareness with customers and improve adoption of XBRL. There were roughly 250 attendees at the event, about half of which were vendors and consultants and the rest financial reporting staff from corporate filers.  Event sponsors included Ernst & Young, SWIFT and Fujitsu.  There were also a number of XBRL technology and service providers exhibiting at the conference.  On Monday Nov. 8th, the XBRL US Steering Committee meetings and Annual Members meeting and reception were held.  At the Annual Members meeting the big news was that current XBRL US President, Mark Bolgiano, is moving to a new position at Howard Hughes Medical Center.  Campbell Pryde, who had led the Taxonomy Development for XBRL US, is taking over as XBRL US President. Other items that were highlighted at the members meeting included: The US GAAP XBRL taxonomy is being used by over 1500 SEC filers and has now been handed over to the FASB to maintain and enhance 16 filer training events were held in 2010 XBRL Global Magazine was launched Corporate Actions proposal was submitted to the SEC with SWIFT in May XBRL Labs for iPhone, XBRL US Consistency Suite launched ISO 2022 Corporate Actions Alignment with XBRL achieved The XBRL Credit Rating taxonomy was accepted Tuesday Nov. 9th included Keynotes, General Sessions, Innovation Workshop for Governments and Securities Professionals, and an Opening Reception.  General sessions included: Lessons Learned from the SEC's rollout of XBRL.  More than 18,000 errors were identified in reviews of filings between June 2009 and September 2010.  Most of these related to negative values being used where they shouldn't have.  Also, the SEC feels there are too many taxonomy extensions being created - mostly in the Cash Flow Statements.  They emphasize using existing elements in the US GAAP taxonomy and advise filers not to  create extensions to improve the visual formatting of XBRL filings. Investors and XBRL - Setting the Standard for Data Quality.  In this panel discussion, the key learning was that CFA's, academics and the financial community are not using XBRL as expected.  The issues raised include the  accuracy and completeness of filings, number of taxonomy extensions, and limited number of tools available to help analyze XBRL data.  Another big issue that was raised is the lack of historic results in XBRL - most analysts need 10 quarters of historic data.  On the positive side, XBRL has the potential to eliminate re-keying of data and errors here and can improve analytic capabilities for financial analysts once more historic data is available and more companies are providing detailed tagging of their filings. A US Roadmap for XBRL Financial Reporting.  This was a panel discussion featuring Jeff Neumann(SEC), Campbell Pryde(XBRL US), and Louis Matherne(FASB).  Key points included the fact that XBRL is currently used by 1500 companies, with 8000 more companies coming in 2011.  XBRL for Mutual Fund Reporting will start in 2011 for 8000 funds, and a Credit Rating Taxonomy has now been submitted for review.  The XBRL tagging/filing process is improving each quarter - more education is helping here.  The FASB is looking at extensions to date, and potential additions to US GAAP taxonomy, while the SEC is evaluating filings for accuracy, consistency in tagging, and tools for analyzing data.  The big news is that the FASB 2011 US GAAP Taxonomy has been completed and reviewed by SEC.  The 2011 US GAAP Taxonomy supports new FASB accounting standards issued since 2009, has new taxonomy elements for certain industries (i.e airlines) and the elimination of 500 concepts.  (meaning they can't be used going forward but are still supported for historical comparison)  The 2011 US GAAP Taxonomy will be available for usage with Q2 2011 SEC filings.  More information about this can be found on the FASB web site.  http://www.fasb.org/home Accounting Firms and XBRL.  This session covered the Role of Audit Firms, which includes awareness and education, validation of XBRL filings, and in-house transition planning.  The main advice provided was that organizations should document XBRL mapping process, perform peer comparisons, and risk assessments on a regular basis. Wednesday Nov. 10th included more Keynotes, General Sessions on Corporate Actions, and XBRL Essentials Workshop Training for corporate filers.  The XBRL Essentials Training included: Getting Started Once you Have the Basics Detailed Footnote Tagging and Handling Tables Quality Control and Trust in the XBRL Process Bringing XBRL In-House:  What are the Options, What should you consider? The US GAAP Financial Reporting Taxonomy - Overview of the 2011 release The XBRL Essentials Training was well-attended with about 80 people.  This included a good overview of the SEC's XBRL mandate, limited liability issue, tagging levels, recommended planning process, internal vs. outsourced approach, and how to manage service providers.  I learned a lot from the session on detailed tagging.  This is the requirement that kicks in during a company's second year of XBRL filing with the SEC and applies to financial statements, footnotes and disclosures (it does not apply to MD&A, executive communications and other information).  The review of the Linkbase model, or dimensional table structure, was very interesting and can be complex to understand.  The key takeaway here is that using dimensional tables in XBRL filings can help limit the number of taxonomy extensions that are required.  The slides from this session are posted on the XBRL US web site. (http://xbrl.us/events/Pages/archive.aspx) For me, the main summary points and takeaways from the XBRL US conference are: XBRL for financial reporting has turned the corner and gone mainstream - with 1500 companies currently using it and 8000 more coming in 2011 The expected value is not being achieved by filers or consumers of XBRL data - this will improve when more companies are filing in XBRL, more history is available, and more software tools are available for analysis (hmm, sounds like an opportunity for Oracle) XBRL is becoming the global standard for all business communications beyond just the financials - i.e. adoption for mutual funds, corporate actions and others planned for the future If you would like to learn more about XBRL and the various training programs, services and software tools that are available check out the XBRL US web site and even better - become a member.  Here's a link:  http://xbrl.us/Pages/default.aspx

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  • Personal Development : Time, Planning , Repairs & Maintenance

    - by Rajesh Pillai
    Personal Development : Time, Planning, Repairs & Maintenance These are just my thoughts, but some you may find something interesting in it. Please think over it. We may know many things, but still we always keeps procrastinating it. I have written this as I have heard many people coming back and saying they don’t have time to do things they like. These are my thoughts buy may be useful to someone else too. Certain things in life needs periodic repairs and maintenance. To cite some examples , your CAR, your HOUSE, your personal laptop/desktop, your health etc. Likewise there are certain other things in professional life that requires repair/ maintenance /or some kind of polishing, so that you always stay on top of it. But they are not always obvious. Some of them are - Improving your communication skills - Increasing your vocabulary - Upgrading your technical skills - Pursuing your hobby - Increasing your knowledge/awareness etc… etc… And then there are certain things that we are always short of…. one is TIME. We all know TIME is one of the most precious things in life and yet we all are very miserable at managing it. Remember you can only manage it and not control it. You can only control which you own or which you create. In theory time is infinite. So, there should be abundant of it. But remember one thing, you know this, it’s not reversible. Once it has elapsed you cannot live it again. Think over it. So, how do find that golden 25th hour every day. To find the 25th hour you need to reflect back on your current daily activities. Analyze them and see where you are spending most of your time and is it really important. Even the 8 hours that you spent in the office, is it spent fruitfully. At the end of the day is the 8 precious hour that you spent was worth it. Just reflect back on your activities. Did you learn something? If yes did you make a point to NOTE IT. If you didn’t NOTED it then was the time you spent really worth it. Just ponder over it. Some calculations of your daily activities where most of the time is spent. Let’s start (in no particular order though) - Sleep (6.5 hours) [Remember you only require 6 good hours of sleep every day]. Some may thing it is 8, but it’s a myth.   o To achive 6 hours of sleep and be in good health you can practice 15 minutes of daily meditation. So effectively you can    round it to 6.5 hours. - Morning chores(2 hours) : Some may need to prepare breakfast and all other things. - Office commuting (avg. to and fro 3 hours) - Office Work (avg 9.5 hours) Total Hours: 21 hours effective time which is spent irrespective of what you do. There may be some variations here and there. Still you have 3 hours EXTRA. Where do these 3 hours go? If you can find it, then you may get that golden 25th hour out of these 3 hours. Let’s discount 2 hours for contingencies, still you have 1 hour with you. If you can’t find it then you are living a direction less life. As you can see, the 25th Hour lies within the 24 hours of the day. It’s upto each one of us to find and make use of it. Now what can you do with that 25th hour i.e. 1 hour extra of your life. Imagine the possibility. Again some calculations 1 hour daily * 30 days = 30 hours every month 30 hours pm * 12 month = 360 hours every year. 360 hours every year seems very promising. Let’s add some contingencies, say, let’s be optimistic and say 50 % contingency. Still you have 180 hours every year. That leaves with 30 minutes every day of extra time. That’s hell a lot of time, if you could manage it. These may sound like a high talk [yes, it is, unless you apply these simple rules and rationalize your everyday living and stop procrastinating]. NOTE: I haven’t taken weekend, holidays and leaves into account. So, that leaves us with a lot of buffer time. You can meet family friends, relatives, other tasks, and yet have these 180 pure hours of joy every year. Do whatever you want to do with it. So, how important is this 180 hours per year to you? Just think over it. You may use it the way you like - 50 hours [pursue your hobby like drawing, crafting, learn dance, learn juggling, learn swimming, travelling hmm.. anything you like doing and you didn’t had time to do it.] - 30 hours you can learn a new programming language or technology (i.e. you can get comfortable with it) - 50 hours [improve existing skills] - 20 hours [improve you communication skill]. Do some light reading. - 30 hours [YOU DECIDE WHAT TO DO]? So, if you had done this for one year you would have learnt a new programming language, upgraded existing skills, improved you communication etc.. If you had done this for two years.. imagine the level of personal development or growth which you may have attained….. If you had done this for three years….. NOW I think I don’t need to mention this… So, you still have TIME, as they say TIME is infinite. So, make judicious use of this precious thing. And never ever comeback saying “I don’t have time”. So, if you are RICH in TIME, everything else will be automatically taken care of, as those things may just be a byproduct of how you spend your time… So, happy TIMING your TIME everyday.

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  • Fun with Declarative Components

    - by [email protected]
    Use case background I have been asked on a number of occasions if our selectOneChoice component could allow random text to be entered, as well as having a list of selections available. Unfortunately, the selectOneChoice component only allows entry via the dropdown selection list and doesn't allow text entry. I was thinking of possible solutions and thought that this might make a good example for using a declarative component.My initial idea My first thought was to use an af:inputText to allow the text entry, and an af:selectOneChoice with mode="compact" for the selections. To get it to layout horizontally, we would want to use an af:panelGroupLayout with layout="horizontal". To get the label for this to line up correctly, we'll need to wrap the af:panelGroupLayout with an af:panelLabelAndMessage. This is the basic structure: <af:panelLabelAndMessage> <af:panelGroupLayout layout="horizontal"> <af:inputText/> <af:selectOneChoice mode="compact"/> </af:panelgroupLayout></af:panelLabelAndMessage> Make it into a declarative component One of the steps to making a declarative component is deciding what attributes we want to be able to specify. To keep this example simple, let's just have: 'label' (the label of our declarative component)'value' (what we want to bind to the value of the input text)'items' (the select items in our dropdown) Here is the initial declarative component code (saved as file "inputTextWithChoice.jsff"): <?xml version='1.0' encoding='UTF-8'?><!-- Copyright (c) 2008, Oracle and/or its affiliates. All rights reserved. --><jsp:root xmlns:jsp="http://java.sun.com/JSP/Page" version="2.1" xmlns:f="http://java.sun.com/jsf/core" xmlns:af="http://xmlns.oracle.com/adf/faces/rich"> <jsp:directive.page contentType="text/html;charset=utf-8"/> <af:componentDef var="attrs" componentVar="comp"> <af:xmlContent> <component xmlns="http://xmlns.oracle.com/adf/faces/rich/component"> <description>Input text with choice component.</description> <attribute> <description>Label</description> <attribute-name>label</attribute-name> <attribute-class>java.lang.String</attribute-class> </attribute> <attribute> <description>Value</description> <attribute-name>value</attribute-name> <attribute-class>java.lang.Object</attribute-class> </attribute> <attribute> <description>Choice Select Items Value</description> <attribute-name>items</attribute-name> <attribute-class>[[Ljavax.faces.model.SelectItem;</attribute-class> </attribute> </component> </af:xmlContent> <af:panelLabelAndMessage id="myPlm" label="#{attrs.label}" for="myIt"> <af:panelGroupLayout id="myPgl" layout="horizontal"> <af:inputText id="myIt" value="#{attrs.value}" partialTriggers="mySoc" label="myIt" simple="true" /> <af:selectOneChoice id="mySoc" label="mySoc" simple="true" mode="compact" value="#{attrs.value}" autoSubmit="true"> <f:selectItems id="mySIs" value="#{attrs.items}" /> </af:selectOneChoice> </af:panelGroupLayout> </af:panelLabelAndMessage> </af:componentDef></jsp:root> By having af:inputText and af:selectOneChoice both have the same value, then (assuming that this passed in as an EL expression) selecting something in the selectOneChoice will update the value in the af:inputText. To use this declarative component in a jspx page: <af:declarativeComponent id="myItwc" viewId="inputTextWithChoice.jsff" label="InputText with Choice" value="#{demoInput.choiceValue}" items="#{demoInput.selectItems}" /> Some problems arise At first glace, this seems to be functioning like we want it to. However, there is a side effect to having the af:inputText and af:selectOneChoice share a value, if one changes, so does the other. The problem here is that when we update the af:inputText to something that doesn't match one of the selections in the af:selectOneChoice, the af:selectOneChoice will set itself to null (since the value doesn't match one of the selections) and the next time the page is submitted, it will submit the null value and the af:inputText will be empty. Oops, we don't want that. Hmm, what to do. Okay, how about if we make sure that the current value is always available in the selection list. But, lets not render it if the value is empty. We also need to add a partialTriggers attribute so that this gets updated when the af:inputText is changed. Plus, we really don't want to select this item so let's disable it. <af:selectOneChoice id="mySoc" partialTriggers="myIt" label="mySoc" simple="true" mode="compact" value="#{attrs.value}" autoSubmit="true"> <af:selectItem id="mySI" label="Selected:#{attrs.value}" value="#{attrs.value}" disabled="true" rendered="#{!empty attrs.value}"/> <af:separator id="mySp" /> <f:selectItems id="mySIs" value="#{attrs.items}" /></af:selectOneChoice> That seems to be working pretty good. One minor issue that we probably can't do anything about is that when you enter something in the inputText and then click on the selectOneChoice, the popup is displayed, but then goes away because it has been replaced via PPR because we told it to with the partialTriggers="myIt". This is not that big a deal, since if you are entering something manually, you probably don't want to select something from the list right afterwards. Making it look like a single component. Now, let's play around a bit with the contentStyle of the af:inputText and the af:selectOneChoice so that the compact icon will layout inside the af:inputText, making it look more like an af:selectManyChoice. We need to add some padding-right to the af;inputText so there is space for the icon. These adjustments were for the Fusion FX skin. <af:inputText id="myIt" partialTriggers="mySoc" autoSubmit="true" contentStyle="padding-right: 15px;" value="#{attrs.value}" label="myIt" simple="true" /><af:selectOneChoice id="mySoc" partialTriggers="myIt" contentStyle="position: relative; top: -2px; left: -19px;" label="mySoc" simple="true" mode="compact" value="#{attrs.value}" autoSubmit="true"> <af:selectItem id="mySI" label="Selected:#{attrs.value}" value="#{attrs.value}" disabled="true" rendered="#{!empty attrs.value}"/> <af:separator id="mySp" /> <f:selectItems id="mySIs" value="#{attrs.items}" /></af:selectOneChoice> There you have it, a declarative component that allows for suggested selections, but also allows arbitrary text to be entered. This could be used for search field, where the 'items' attribute could be populated with popular searches. Lines of java code written: 0

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  • Day 3 - XNA: Hacking around with images

    - by dapostolov
    Yay! Today I'm going to get into some code! My mind has been on this all day! I find it amusing how I practice, daily, to be "in the moment" or "present" and the excitement and anticipation of this project seems to snatch it away from me frequently. WELL!!! (Shakes Excitedly) Let's do this =)! Let's code! For these next few days it is my intention to better understand image rendering using XNA; after said prototypes are complete I should (fingers crossed) be able to dive into my game code using the design document I hammered out the other night. On a personal note, I think the toughest thing right now is finding the time to do this project. Each night, after my little ones go to bed I can only really afford a couple hours of work on this project. However, I hope to utilise this time as best as I can because this is the first time in a while I've found a project that I've been passionate about. A friend recently asked me if I intend to go 3D or extend the game design. Yes. For now I'm keeping it simple. Lastly, just as a note, as I was doing some further research into image rendering this morning I came across some other XNA content and lessons learned. I believe this content could have probably been posted in the first couple of posts, however, I will share the new content as I learn it at the end of each day. Maybe I'll take some time later to fix the posts but for now Installation and Deployment - Lessons Learned I had installed the XNA studio  (Day 1) and the site instructions were pretty easy to follow. However, I had a small difficulty with my development environment. You see, I run a virtual desktop development environment. Even though I was able to code and compile all the tutorials the game failed to run...because I lacked a 3D capable card; it was not detected on the virtual box... First Lesson: The XNA runtime needs to "see" the 3D card! No sweat, Il copied the files over to my parent box and executed the program. ERROR. Hmm... Second Lesson (which I should have probably known but I let the excitement get the better of me): you need the XNA runtime on the client PC to run the game, oh, and don't forget the .Net Runtime! Sprite, it ain't just a Soft Drink... With these prototypes I intend to understand and perform the following tasks. learn game development terminology how to place and position (rotate) a static image on the screen how to layer static images on the screen understand image scaling can we reuse images? understand how framerate is handled in XNA how to display text , basic shapes, and colors on the screen how to interact with an image (collision of user input?) how to animate an image and understand basic animation techniques how to detect colliding images or screen edges how to manipulate the image, lets say colors, stretching how to focus on a segment of an image...like only displaying a frame on a film reel what's the best way to manage images (compression, storage, location, prevent artwork theft, etc.) Well, let's start with this "prototype" task list for now...Today, let's get an image on the screen and maybe I can mark a few of the tasks as completed... C# Prototype1 New Visual Studio Project Select the XNA Game Studio 3.1 Project Type Select the Windows Game 3.1 Template Type Prototype1 in the Name textbox provided Press OK. At this point code has auto-magically been created. Feel free to press the F5 key to run your first XNA program. You should have a blue screen infront of you. Without getting into the nitty gritty right, the code that was generated basically creates some basic code to clear the window content with the lovely CornFlowerBlue color. Something to notice, when you move your mouse into the window...nothing. ooooo spoooky. Let's put an image on that screen! Step A - Get an Image into the solution Under "Content" in your Solution Explorer, right click and add a new folder and name it "Sprites". Copy a small image in there; I copied a "Royalty Free" wizard hat from a quick google search and named it wizards_hat.jpg (rightfully so!) Step B - Add the sprite and position fields Now, open/edit  Game1.cs Locate the following line:  SpriteBatch spriteBatch; Under this line type the following:         SpriteBatch spriteBatch; // the line you are looking for...         Texture2D sprite;         Vector2 position; Step C - Load the image asset Locate the "Load Content" Method and duplicate the following:             protected override void LoadContent()         {             spriteBatch = new SpriteBatch(GraphicsDevice);             // your image name goes here...             sprite = Content.Load<Texture2D>("Sprites\\wizards_hat");             position = new Vector2(200, 100);             base.LoadContent();         } Step D - Draw the image Locate the "Draw" Method and duplicate the following:        protected override void Draw(GameTime gameTime)         {             GraphicsDevice.Clear(Color.CornflowerBlue);             spriteBatch.Begin(SpriteBlendMode.AlphaBlend);             spriteBatch.Draw(sprite, position, Color.White);             spriteBatch.End();             base.Draw(gameTime);         }  Step E - Compile and Run Engage! (F5) - Debug! Your image should now display on a cornflowerblue window about 200 pixels from the left and 100 pixels from the top. Awesome! =) Pretty cool how we only coded a few lines to display an image, but believe me, there is plenty going on behind the scenes. However, for now, I'm going to call it a night here. Blogging all this progress certainly takes time... However, tomorrow night I'm going to detail what we just did, plus start checking off points on that list! I'm wondering right now if I should add pictures / code to this post...let me know if you want them =) Best Regards, D.

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  • I Clobbered a Leopard with a Window Last Night

    - by D'Arcy Lussier
    I’ve had my 15” Mac Book Pro for a little over a year now, and its hands-down the best laptop I’ve ever owned…hardware wise. And I tried, I really really tried, to like OSX. I even bought Parallels so I could run Windows 7 and all my development tools while still trying to live in an OSX world. But in the end, I missed Windows too much. There were just too many shortcomings with OSX that kept me from being productive. For one thing, Office for Mac is *not* Office for Windows. The applications are written by different teams, and Excel on the Mac is just different enough to be painful. The VM experience was adequate, but my MBP would heat up like crazy when running it and the experience trying to get Windows apps to interact with an OSX file system was awkward. And I found I was in the VM more than I thought I’d be. iMovie is not as easy to use for doing simple movie editing as Windows Movie Maker. There’s no free blog editing software for OSX that’s on par with Windows Live Writer. And really, all I was using OSX for was Twitter (which I can use a Windows client for) and web browsing (also something Windows can provide obviously). So I had to ask myself – why am I forcing myself to use an operating system I don’t like, on a laptop that can support Windows 7? And so I paved my MBP and am happily running Windows 7 on it…and its fantastic! All the good stuff with the hardware is still there with the goodness of Win 7. Happy happy. I did run into some snags doing this though, and that’s really what this blog post is about – things to be aware of if you want to install Win 7 directly on your MBP metal. First, Ensure You Have Your Original Mac Install Disk This was a warning my buddy Dylan, who’s been running Win 7 on his MBP for a while now, gave me early on. The reason you need that original disk is that the hardware drivers you need are all located there. Apparently you can’t easily download them, so make sure you have them ahead of time. Second, Forget BootCamp The only reason you need BootCamp is if you still want the option to boot into OSX. If you don’t, then you don’t need BootCamp. In fact, you don’t even need BootCamp to install Win 7. What you *will* need though is a DVD with Win 7 burnt on it. Apple doesn’t support bootable USB drives. Well, actually they do for Mac Book Airs which don’t come with optical drives…but to get it working you’ll need to edit a system file of BootCamp so your make of MBP is included in an XML document, and even then you *still* are using BootCamp meaning you’ll be making an OSX partition. So don’t worry about BootCamp, just burn a Windows 7 disc, put it into the DVD drive, and restart your MBP. Third, Know The Secret Commands So after putting in the Windows 7 DVD and restarting your MBP, you’ll want to hold down the ‘C’ key during boot up. This tells the MBP that it should boot from the DVD drive instead of the hard drive. Interestingly, it appears you don’t have to do this if its the Mac OSX install disc (more on that in a second), but regardless – hold down C and Windows will start the install process. Next up is the partition process. You’ll notice that there’s a partition called ETI or something like that. This has to do with the drive format that Apple uses and how they partition their system drives. What I did – I blew it away! At first I didn’t, but I was told I couldn’t install Windows on the remaining space due to the different drive format. Blowing away the ETI partition (and all other partitions) allowed me to continue the Windows install. *REMEMBER –  No warranty is provided or implied, just telling you what I did and how I got it to work. Ok, so now Windows is installed and I’m rebooting. Everything looks good, but I need drivers! So I put in the OSX install DVD and run the BootCamp assistant which installs all the Windows drivers I need. Fantastic! Oh, I need to restart – no problem. OH NO, PROBLEM! I left the OSX install DVD in the drive and now the MBP wants to boot from the drive and install OSX! I’m not holding down the C key, what the heck?! Ok, well there must be a way to eject this disk…hmm…no physical button on the side…the eject button doesn’t seem to work on the keyboard…no little pin hole to insert something to force the disc out…well what the…?! It turns out, if you want to eject a disc at boot up, you need (and I kid you not) to plug a mouse into the laptop and hold down the right-click button while its booting. This ejected the disc for me. Seriously. Finally, Things You Should Be Aware Of Once you have Windows up and running there’s a few things you need to be aware of, mainly new keyboard shortcuts. For instance, on the Mac keyboard there is no Home, End, PageUp or PageDown. There’s also no obvious way to do something like select large amounts of text (like you would by holding Shift-Home at the end of a line of text for instance). So here’s some shortcuts you need to know: Home – fn + left arrow End – fn + right arrow Select a line of text as you would with the Home key – Shift + fn + left arrow Select a line of text as you would with the End key – Shift + fn + right arrow Page Up – fn + up arrow Page Down – fn + down arrow Also, you’ll notice that the awesome Mac track pad doesn’t respond to taps as clicks. No fear, this is just a setting that needs to be altered in the BootCamp control panel (that controls the Mac Hardware-specific settings within Windows, you can access it easily from the system tray icon) One other thing, battery life seems a bit lower than with OSX, but then again I’m also doing more than Twitter or web browsing on this thing now. Conclusion My laptop runs awesome now that I have Windows 7 on there. It’s obviously up to individual taste, but for me I just didn’t see benefits to living in an OSX world when everything I needed lived in Windows. And also, I finally am back to an operating system that doesn’t require me to eject a USB drive before physically removing it! It’s 2012 folks, how has this not been fixed?! D

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