Search Results

Search found 568 results on 23 pages for 'designated initializer'.

Page 19/23 | < Previous Page | 15 16 17 18 19 20 21 22 23  | Next Page >

  • Inserting default "admin" user into database during Rails App startup

    - by gbc
    I'm building my first real rails application for a little in-house task. Most of the application tasks require authentication/authorization. The first time the app starts up (or starts with a wiped db), I'd like the process to be: User logs into the admin panel using "admin" & "admin" for authentication info. User navigates to admin credentials editing page and changes name and password to something safer so that "admin" & "admin" is no longer a valid login. To achieve this result, I'd like to stuff a default username & password combination into the database on if the application starts up and detects that there are no user credentials in the 'users' table. For example: if User.count == 0 User.create(:name => "admin", :password => "admin") end However, I'm unsure where to place that code. I tried adding an initializer script in the config/initializers, but the error I received appeared to indicate that the model classes weren't yet loaded into the application. So I'm curious to know at what point I can hook into the application startup cycle and insert data into the database through ActiveRecord before requests are dispatched.

    Read the article

  • template warnings and error help, (gcc)

    - by sil3nt
    Hi there, I'm working on an container class template (for int,bool,strings etc), and I've been stuck with this error cont.h:56: error: expected initializer before '&' token for this section template <typename T> const Container & Container<T>::operator=(const Container<T> & rightCont){ what exactly have I done wrong there?. Also not sure what this warning message means. cont.h:13: warning: friend declaration `bool operator==(const Container<T>&, const Container<T>&)' declares a non-template function cont.h:13: warning: (if this is not what you intended, make sure the function template has already been declared and add <> after the function name here) -Wno-non-template-friend disables this warning at this position template <typename T> class Container{ friend bool operator==(const Container<T> &rhs,const Container<T> &lhs); public:

    Read the article

  • threading in c#

    - by I__
    i am using this code: private void Form1_Load(object sender, EventArgs e) { } private void serialPort1_DataReceived(object sender, System.IO.Ports.SerialDataReceivedEventArgs e) { string response = serialPort1.ReadLine(); this.BeginInvoke(new MethodInvoker( () => textBox1.AppendText(response + "\r\n") )); } ThreadStart myThreadDelegate = new ThreadStart(ThreadWork.DoWork); Thread myThread = new Thread(myThreadDelegate); myThread.Start(); but am getting lots of errors: Error 2 The type or namespace name 'ThreadStart' could not be found (are you missing a using directive or an assembly reference?) C:\Users\alexluvsdanielle\AppData\Local\Temporary Projects\WindowsFormsApplication1\Form1.cs 31 44 WindowsFormsApplication1 Error 3 The name 'ThreadWork' does not exist in the current context C:\Users\alexluvsdanielle\AppData\Local\Temporary Projects\WindowsFormsApplication1\Form1.cs 31 56 WindowsFormsApplication1 Error 4 The type or namespace name 'Thread' could not be found (are you missing a using directive or an assembly reference?) C:\Users\alexluvsdanielle\AppData\Local\Temporary Projects\WindowsFormsApplication1\Form1.cs 32 31 WindowsFormsApplication1 Error 5 A field initializer cannot reference the non-static field, method, or property 'WindowsFormsApplication1.Form1.myThreadDelegate' C:\Users\alexluvsdanielle\AppData\Local\Temporary Projects\WindowsFormsApplication1\Form1.cs 32 38 WindowsFormsApplication1 what am i doing wrong?

    Read the article

  • C++ using typedefs in non-inline functions

    - by ArunSaha
    I have a class like this template< typename T > class vector { public: typedef T & reference; typedef T const & const_reference; typedef size_t size_type; const_reference at( size_t ) const; reference at( size_t ); and later in the same file template< typename T > typename vector<T>::const_reference // Line X vector<T>::at( size_type i ) const { rangecheck(); return elems_[ i ]; } template< typename T > reference // Line Y vector<T>::at( size_type i ) { rangecheck(); return elems_[ i ]; } Line X compiles fine but Line Y does not compile. The error message from g++ (version 4.4.1) is: foo.h:Y: error: expected initializer before 'vector' From this I gather that, if I want to have non-inline functions then I have to fully qualify the typedef name as in Line X. (Note that, there is no problem for size_type.) However, at least to me, Line X looks clumsy. Is there any alternative approach?

    Read the article

  • Approach For Syncing One SharePoint List With One or More SharePoint Lists

    - by plattnum
    What would be the best approach or strategy for configuring, customizing or developing in SharePoint a solution that allows me to keep one or more SharePoint lists in sync with a SharePoint list I have designated as a master or parent list. I would like to be able to create a master/parent list of some information that can be extended or used by different parts of the organization without them being able to CRUD any items on the actual columns of the master list. (I have seen some commercial web parts that offer column security on SharePoint lists and although that’s one way of potentially meeting my needs I would like to explore other options.) Scenario: I have a list called FOO: FOO Title Description I would like to create a new list BAR based off of FOO (BAR is managed by sub-organization that doesn't have access to FOO List): BAR FOO.Title (Read-Only) FOO.Description (Read-Only) NewColumn1 NewColumn2 Actions: Create- If a new item is entered in FOO I would like the new item added to BAR. Read - N/A Update - If the title or description is changed in FOO I would like it changed in BAR. Delete- No Deletes in the scenario. (Deletes are handled by the business with status column.) Templates with content extraction offer me this but it’s a one time shot at list creation. Just not sure what the best approach or strategy would be for this in MOSS 2007. Thanks!

    Read the article

  • How to add a has_many association on all models

    - by joshsz
    Right now I have an initializer that does this: ActiveRecord::Base.send :has_many, :notes, :as => :notable ActiveRecord::Base.send :accepts_nested_attributes_for, :notes It builds the association just fine, except when I load a view that uses it, the second load gives me: can't dup NilClass from: /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2184:in `dup' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2184:in `scoped_methods' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2188:in `current_scoped_methods' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2171:in `scoped?' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2439:in `send' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/base.rb:2439:in `initialize' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/reflection.rb:162:in `new' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/reflection.rb:162:in `build_association' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/associations/association_collection.rb:423:in `build_record' /usr/lib/ruby/gems/1.8/gems/activerecord-2.3.5/lib/active_record/associations/association_collection.rb:102:in `build' (my app)/controllers/manifests_controller.rb:21:in `show' Any ideas? Am I doing this the wrong way? Interestingly if I move the association onto just the model I'm working with at the moment, I don't get this error. I figure I must be building the global association incorrectly.

    Read the article

  • How do I bind to a custom view in Cocoa using Xcode 4?

    - by Newt
    I'm a beginner when it comes to writing Mac apps and working with Cocoa, so please forgive my ignorance. I'm looking to create a custom view, that exposes some properties, which I can then bind to an NSObjectController. Since it's a custom view, the Bindings Inspector obviously doesn't list any of the properties I've added to the view that I can then bind to using Interface Builder. After turning to the Stackoverflow/Google for help, I've stumbled across a couple of possible solutions, but neither seem to be quite right for my situation. The first suggested creating an IBPlugin, which would then mean my bindings would be available in the Bindings Inspector. I could then bind the view to the controller using IB. Apparently IBPlugins aren't supported in Xcode 4, so that one's out the window. I'm also assuming (maybe wrongly) that IBPlugins are no longer supported because there's a better way of doing such things these days? The second option was to bind the controller to the view programmatically. I'm a bit confused as to exactly how I would achieve this. Would it require subclassing NSObjectController so I can add the calls to bind to the view? Would I need to add anything to the view to support this? Some examples I've seen say you'd need to override the bind method, and others say you don't. Also, I've noticed that some example custom views call [self exposeBinding:@"bindingName"] in the initializer. From what I gather from various sources, this is something that's related to IBPlugins and isn't something I need to do if I'm not using them. Is that correct? I've found a post on Stackoverflow here which seems to discuss something very similar to my problem, but there wasn't any clear winner as to the best answer. The last comment by noa on 12th Sept seems interesting, although they mention you should be calling exposeBinding:. Is this comment along the right track? Is the call to exposeBinding really necessary? Apologies for any dumb questions. Any help greatly appreciated.

    Read the article

  • Void* array casting to float, int32, int16, etc.

    - by Griffin
    Hey guys, I've got an array of PCM data, it could be 16 bit, 24 bit packed, 32 bit, etc.. It could be signed, or unsigned, and it could be 32 or 64 bit floating point. It is currently stored as a "void**" matrix, indexed by channel, then by frame. The goal is to allow my library to take in any PCM format and buffer it, without requiring manipulation of the data to fit a designated structure. If the A/D converter spits out 24 bit packed arrays of interleaved PCM, I need to accept it gracefully. I also need to support 16 bit non interleaved, as well as any permutation of the above formats. I know the bit depth and other information at runtime, and I'm trying to code efficiently while not duplicating code. What I need is an effective way to cast the matrix, put PCM data into the matrix, and then pull it out later. I can cast the matrix to int32_t, or int16_t for the 32 and 16 bit signed PCM respectively, I'll probably have to store the 24 bit PCM in an int32_t for 32 bit, 8 bit byte systems as well. Can anyone recommend a good way to put data into this array, and pull it out later? I'd like to avoid large sections of code which look like: switch( mFormat ) { case 1: // unsigned 8 bit for( int i = 0; i < mChannels; i++ ) framesArray = (uint8_t*)pcm[i]; break; case 2: // signed 8 bit for( int i = 0; i < mChannels; i++ ) framesArray = (int8_t*)pcm[i]; break; case 3: // unsigned 16 bit ... Limitations: I'm working in C/C++, no templates, no RTTI, no STL. Think embedded. Things get trickier when I have to port this to a DSP with 16 bit bytes. Does anybody have any useful macros they might be willing to share? Thanks, -Griff

    Read the article

  • Update Options on Existing jQuery Object

    - by Vince Kronlein
    I'm using Bootstrap and DataTables in my app and I have a default initializer for tables based on class. I can just add the class data-table to the table and it gets instantiated with the default values I want. I'd like to know how to change or update specific options based on a specific table. if ($.fn.dataTable) { $('.data-table').dataTable( { sDom: "R<'row'<'span6'l><'span6'f>r>t<'row'<'span6'i><'span6'p>>", sPaginationType: "bootstrap", oLanguage: { "sLengthMenu": "_MENU_ &nbsp; records per page" }, aoColumnDefs: [ { "bSortable": false, "aTargets": [ 0 ] } ] }); } All my data tables have a checkbox in the first column so the above removal of sorting works for all of them. But I'd like to be able to update the aoColumnDefs on a table by table basis so I can add other columns that I don't want sorted. So let's say I have a table: $('#member-list'), how do I access this object and update it's datatables options in jQuery? I can't find any reference or help anywhere. Thanks a lot! -V

    Read the article

  • What should I do from here?

    - by Sunscreen
    Hi all, First of all, the site rocks. You can ask and get specific answers, mainly, for programming issues. This question is more generic. I studied Physics for my bachelors and Digital Image Processing for my masters, ended on September 2001. From then on I started working as a developer and software analyst. I worked, and working, witn C, C++, AIX OS, XP OS, MFC 4.21. I also did some data translations from EDIFACT to XML and viceversa. I trained users for the applications that I was running, I created documents (detailed design docs mainly), though most of the time I wrote, and I still write, code. Recently I applied for the best greek, graduate university for my MBA and they accepted me, starting on Jan 2011. I am a developer with no specific insight with the languages I work. I can be very productive with some subsets of the languages that the companies I worked for use, though this is a limited thing for a developer. If I get my MBA I can be a semi-businees analyst ot consultant, as I am now a semi-developer. The problem is that I can do some but not all in a designated working area. What should I do from here? Should I get my MBA and look forswitching industries? Should I read and excersise myself with new languages and frameworks? Should I be more focussed to the deligations from my current job (currently I work with MFC)? Just for the note, I am 32 and I feel I am wasting my time. I am not getting the best that I can get from is current position (and I work here for 3+ years). Thanks all, Sun

    Read the article

  • How to make Spring load a JDBC Driver BEFORE initializing Hibernate's SessionFactory?

    - by Bill_BsB
    I'm developing a Spring(2.5.6)+Hibernate(3.2.6) web application to connect to a custom database. For that I have custom JDBC Driver and Hibernate Dialect. I know for sure that these custom classes work (hard coded stuff on my unit tests). The problem, I guess, is with the order on which things get loaded by Spring. Basically: Custom Database initializes Spring load beans from web.xml Spring loads ServletBeans(applicationContext.xml) Hibernate kicks in: shows version and all the properties correctly loaded. Hibernate's HbmBinder runs (maps all my classes) LocalSessionFactoryBean - Building new Hibernate SessionFactory DriverManagerConnectionProvider - using driver: MyCustomJDBCDriver at CustomDBURL I get a SQLException: No suitable driver found for CustomDBURL Hibernate loads the Custom Dialect My CustomJDBCDriver finally gets registered with DriverManager (log messages) SettingsFactory runs SchemaExport runs (hbm2ddl) I get a SQLException: No suitable driver found for CustomDBURL (again?!) Application get successfully deployed but there are no tables on my custom Database. Things that I tried so far: Different techniques for passing hibernate properties: embedded in the 'sessionFactory' bean, loaded from a hibernate.properties file. Nothing worked but I didn't try with hibernate.cfg.xml file neither with a dataSource bean yet. MyCustomJDBCDriver has a static initializer block that registers it self with the DriverManager. Tried different combinations of lazy initializing (lazy-init="true") of the Spring beans but nothing worked. My custom JDBC driver should be the first thing to be loaded - not sure if by Spring but...! Can anyone give me a solution for this or maybe a hint for what else I could try? I can provide more details (huge stack traces for instance) if that helps. Thanks in advance.

    Read the article

  • Using an initializer_list on a map of vectors

    - by Hooked
    I've been trying to initialize a map of <ints, vector<ints> > using the new 0X standard, but I cannot seem to get the syntax correct. I'd like to make a map with a single entry with key:value = 1:<3,4 #include <initializer_list> #include <map> #include <vector> using namespace std; map<int, vector<int> > A = {1,{3,4}}; .... It dies with the following error using gcc 4.4.3: error: no matching function for call to std::map<int,std::vector<int,std::allocator<int> >,std::less<int>,std::allocator<std::pair<const int,std::vector<int,std::allocator<int> > > > >::map(<brace-enclosed initializer list>) Edit Following the suggestion by Cogwheel and adding the extra brace it now compiles with a warning that can be gotten rid of using the -fno-deduce-init-list flag. Is there any danger in doing so?

    Read the article

  • C++: Copy contructor: Use Getters or access member vars directly?

    - by cbrulak
    Have a simple container class: public Container { public: Container() {} Container(const Container& cont) //option 1 { SetMyString(cont.GetMyString()); } //OR Container(const Container& cont) //option 2 { m_str1 = cont.m_str1; } public string GetMyString() { return m_str1;} public void SetMyString(string str) { m_str1 = str;} private: string m_str1; } So, would you recommend this method or accessing the member variables directly? In the example, all code is inline, but in our real code there is no inline code. Update (29 Sept 09): Some of these answers are well written however they seem to get missing the point of this question: this is simple contrived example to discuss using getters/setters vs variables initializer lists or private validator functions are not really part of this question. I'm wondering if either design will make the code easier to maintain and expand. Some ppl are focusing on the string in this example however it is just an example, imagine it is a different object instead. I'm not concerned about performance. we're not programming on the PDP-11

    Read the article

  • Tableview not drilling down

    - by JoshD
    I have a Tableview which is loaded with a dummy list of exercises. I need to drill to a different page. My didSelectRow is being called, but not implementing the method. The app is a tab bar navigation app that loads a UITableView. #import <UIKit/UIKit.h> @interface Schedule : UIViewController <UITableViewDelegate, UITableViewDataSource> { NSArray *mySchedule; } @end #import "Schedule.h" #import "AnotherViewController.h" @implementation Schedule - (NSInteger)tableView:(UITableView *)tableView numberOfRowsInSection:(NSInteger)section{ return mySchedule.count; } - (UITableViewCell *)tableView:(UITableView *)tableView cellForRowAtIndexPath:(NSIndexPath *)indexPath{ //create a cell UITableViewCell *cell = [[UITableViewCell alloc] initWithStyle:UITableViewCellStyleDefault reuseIdentifier:@"cell"]; //fill it with contents cell.textLabel.text = [mySchedule objectAtIndex:indexPath.row]; //return it return cell; } - (void)tableView:(UITableView *)tableView didSelectRowAtIndexPath:(NSIndexPath *)indexPath{ AnotherViewController *anotherViewController = [[AnotherViewController alloc] initWithNibName:@"AnotherViewController" bundle:nil]; [self.navigationController pushViewController:anotherViewController animated:YES]; [anotherViewController release]; } /* // The designated initializer. Override if you create the controller programmatically and want to perform customization that is not appropriate for viewDidLoad. - (id)initWithNibName:(NSString *)nibNameOrNil bundle:(NSBundle *)nibBundleOrNil { if (self = [super initWithNibName:nibNameOrNil bundle:nibBundleOrNil]) { // Custom initialization } return self; } */ // Implement viewDidLoad to do additional setup after loading the view, typically from a nib. - (void)viewDidLoad { NSString *myFile = [[NSBundle mainBundle] pathForResource:@"exercise" ofType:@"plist"]; mySchedule = [[NSArray alloc] initWithContentsOfFile:myFile]; [super viewDidLoad]; }

    Read the article

  • PHP, MySQL, Memcache / Ajax Scaling Problem

    - by Jeff Andersen
    I'm building a ajax tic tac toe game in PHP/MySQL. The premise of the game is to be able to share a url like mygame.com/123 with your friends and you play multiple simultaneous games. The way I have it set up is that a file (reload.php) is being called every 3 seconds while the user is viewing their game board space. This reload.php builds their game boards and the output (html) replaces their current game board (thus showing games in which it is their turn) Initially I built it entirely with PHP/MySQL and had zero caching. A friend gave me a suggestion to try doing all of the temporary/quick read information through memcache (storing moves, and ID matchups) and then building the game boards from this information. My issue is that, both solutions encounter a wall when there is roughly 30-40 active users with roughly 40-50 games running. It is running on a VPS from VPS.net with 2 nodes. (Dedicated CPU: 1.2GHz, RAM: 752MB) Each call to reload.php peforms 3 selects and 2 insert queries. The size of the data being pulled is negligible. The same actions happen on index.php to build the boards for the initial visit. Now that the backstory is done, my question is: Would there be a bottleneck in that each user is polling the same file every 3 seconds to rebuild their gameboards, and that all users are sitting on index.php from which the AJAX calls are made within the HTML. If so, is it possible to spread the users' calls out over a set of files designated to building the game boards (eg. reload1.php 2, 3 etc) and direct users to the appropriate file. Would this relieve the pressure? A long winded explanation; however, I didn't have anywhere else to ask. Thanks very much for any insight.

    Read the article

  • C++ reference variable again!!!

    - by kumar_m_kiran
    Hi All, I think most would be surprised about the topic again, However I am referring to a book "C++ Common Knowledge: Essential Intermediate Programming" written by "Stephen C. Dewhurst". In the book, he quotes a particular sentence (in section under Item 5. References Are Aliases, Not Pointers), which is as below A reference is an alias for an object that already exists prior to the initialization of the reference. Once a reference is initialized to refer to a particular object, it cannot later be made to refer to a different object; a reference is bound to its initializer for its whole lifetime Can anyone please explain the context of "cannot later be made to refer to a different object" Below code works for me, #include <iostream> using namespace std; int main(int argc, char *argv[]) { int i = 100; int& ref = i; cout<<ref<<endl; int k = 2000; ref = k; cout<<ref<<endl; return 0; } Here I am referring the variable ref to both i and j variable. And the code works perfectly fine. Am I missing something? I have used SUSE10 64bit linux for testing my sample program. Thanks for your input in advance.

    Read the article

  • Isotope.js help: Changing item image after sorting

    - by user3643081
    This is a general question on how to go about building a project I have in mind, and the best way to set off on the right foot. I am fairly new to JS, please be gentle. I want to use isotope.js (or a similar script) to display a page with multiple items (about 30 different plants found in a garden) and the ability to sort them by seasons of the year + "what is most beautiful now" + and "view all" (a total of 6 categories) . On load, or when sorted by either "what is beautiful now" or "view all", I need each item to reflect the image of the current season we are in. When sorted by season, I need those "current" images to switch over to a designated seasonal image of that plant. Therefore, each sortable item will ultimately have 4 different versions with 4 different images in the background ready to surface when plants are sorted. (perhaps 5 if it makes more sense to have a "current" version besides the 4 seasonal versions.) My question: what approach can I take to achieve this effect in a manageable way? Can isotope apply a class to items sorted? Assuming it can: Should each item have 4 inline images, each with a css class, that I then control by using display:inline; and display:none; properties from my stylesheets? (I worry that this approach would significantly increase load times) Would it make more sense to create a blank dummy div who's background I control similarly to the example above -relying mostly on CSS. Or is there some other way involving JS I am overlooking? Any help would be appreciated. Examples of what you suggest would be immensely helpful.

    Read the article

  • Google Drive API invalid_grant after removing access

    - by Sparafusile
    I have been writing a desktop application that uses the Google Drive API v2. I have the following code: var credential = GoogleWebAuthorizationBroker.AuthorizeAsync ( new ClientSecrets { ClientId = ClientID, ClientSecret = ClientSecret }, new[] { DriveService.Scope.Drive }, "user", CancellationToken.None ) .Result; this.Service = new DriveService( new BaseClientService.Initializer() { HttpClientInitializer = credential, ApplicationName = "My Test App", } ); var request = this.Service.Files.List(); request.Q = "title = 'foo' and trashed = false"; var result = request.Execute(); The first time I ran this code it opened a browser and asked me to grant permissions to the App, which I did. Everything worked successfully until I realized I was using the wrong Google account. At that point I logged into the wrong Google account and revoked access to my App. Now, whenever I run the same code it throws an exception: Error:"invalid_grant", Description:"", Uri:"" When I examine the service and request objects, it looks like the oauth_token isn't getting created any more. I know what I did to mess things up, but I can't figure out how to correct it so I can use a different Google account for testing. What do I need to do?

    Read the article

  • What is a truly empty std::vector in C++?

    - by RyanG
    I've got a two vectors in class A that contain other class objects B and C. I know exactly how many elements these vectors are supposed to hold at maximum. In the initializer list of class A's constructor, I initialize these vectors to their max sizes (constants). If I understand this correctly, I now have a vector of objects of class B that have been initialized using their default constructor. Right? When I wrote this code, I thought this was the only way to deal with things. However, I've since learned about std::vector.reserve() and I'd like to achieve something different. I'd like to allocate memory for these vectors to grow as large as possible because adding to them is controlled by user-input, so I don't want frequent resizings. However, I iterate through this vector many, many times per second and I only currently work on objects I've flagged as "active". To have to check a boolean member of class B/C on ever iteration is silly. I don't want these objects to even BE there for my iterators to see when I run through this list. Is reserving the max space ahead of time and using push_back to add a new object to the vector a solution to this?

    Read the article

  • The best way to predict performance without actually porting the code?

    - by ardiyu07
    I believe there are people with the same experience with me, where he/she must give a (estimated) performance report of porting a program from sequential to parallel with some designated multicore hardwares, with a very few amount of time given. For instance, if a 10K LoC sequential program was given and executes on Intel i7-3770k (not vectorized) in 100 ms, how long would it take to run if one parallelizes the code to a Tesla C2075 with NVIDIA CUDA, given that all kinds of parallelizing optimization techniques were done? (but you're only given 2-4 days to report the performance? assume that you didn't know the algorithm at all. Or perhaps it'd be safer if we just assume that it's an impossible situation to finish the job) Therefore, I'm wondering, what most likely be the fastest way to give such performance report? Is it safe to calculate solely by the hardware's capability, such as GFLOPs peak and memory bandwidth rate? Is there a mathematical way to calculate it? If there is, please prove your method with the corresponding problem description and the algorithm, and also the target hardwares' specifications. Or perhaps there already exists such tool to (roughly) estimate code porting? (Please don't the answer: 'kill yourself is the fastest way.')

    Read the article

  • Should I re-use UI elements across view controllers?

    - by Endemic
    In the iPhone app I'm currently working on, I'd like two navigation controllers (I'll call them A and B) to have toolbars that are identical in appearance and function. The toolbar in question will look like this: [(button) (flexible-space) (label)] For posterity's sake, the label is actually a UIBarButtonItem with a custom view. My design requires that A always appear directly before B on the navigation stack, so B will never be loaded without A having been loaded. Given this layout, I started wondering, "Is it worth it to re-use A's toolbar items in B's toolbar?" As I see it, my options are: 1. Don't worry about re-use, create the toolbar items twice 2. Create the toolbar items in A and pass them to B in a custom initializer 3. Use some more obscure method that I haven't thought of to hold the toolbar constant when pushing a view controller As far as I can see, option 1 may violate DRY, but guarantees that there won't be any confusion on the off chance that (for example) the button may be required to perform two different (no matter how similar) functions for either view controller in future versions of the app. Were that to happen, options 2 or 3 would require the target-action of the button to change when B is loaded and unloaded. Even if the button were never required to perform different functions, I'm not sure what its proper target would be under option 2. All in all, it's not a huge problem, even if I have to go with option 1. I'm probably overthinking this anyway, trying to apply the dependency injection pattern where it's not appropriate. I just want to know the best practice should this situation arise in a more extreme form, like if a long chain of view controllers need to use identical (in appearance and function) UI elements.

    Read the article

  • Windows 8 with LiveID login authenticates as Guest to remote SQl Server

    - by Tim Long
    I have a network where several users are using Office Accounting 2009 in multi-user client/server mode. OA is built on SQL Server. One PC acts as the 'server' and has the SQl Server instance, the others have only the application installed and no SQL instance, all of the apps connect remotely to the SQL instance on the 'server'. I'm using the term 'server' loosely here, it is just a normal workstation that happens to be designated as the server and runs the SQL instance. There is no NT domain, all user accounts are local accounts. The way that OA works in multi-user mode is that each user is required to have a local account with the same username and password on both the client and 'server' PCs. This has been working well, no along comes Windows 8. I use my 'Microsoft Account' aka LiveID to log into Windows 8. Office Accounting runs fine and attempts to connect to the database, but fails, 'you do not have permission to perform this operation'. In the SQL logs, I get this error: 2012-10-28 17:54:01.32 Logon Error: 18456, Severity: 14, State: 11. 2012-10-28 17:54:01.32 Logon Login failed for user 'SERVER\Guest'. Reason: Token-based server access validation failed with an infrastructure SERVER is the hostname of the server. So it seems to be authenticating as 'Guest'?? To verify this, I enabled the Guest account on the 'server' PC and then added Guest as an allowed user within Office Accounting (this simply creates the user in SQL and gives it an appropriate database role). Sure enough, My Windows 8 PC was then able to connect to the database when using Office Accounting. Clearly, having users authenticate as 'Guest' stinks from a security and auditing standpoint. So what I need are some ideas for how to work around this. I've tried switching the Windows 8 PC to a 'local account' and that works too, but requires giving up significant functionality on the Windows 8 PC. What I really need is a way to force the Windows 8 PC to use a specific set of credentials when connecting to the remote SQL instance. Office Accounting takes the logged in username, which is my LiveID and doesn't correspond to any Windows user name. Anyone solved this issue?

    Read the article

  • ISC DHCP - Force clients to get a new IP address, instead of the being re-issued their previous lease's IP

    - by kce
    We are in the middle of a migration of our DHCP and DNS services from a Debian-based server to a Windows Server 2008 R2 implementation. The Debian server is running isc-dhcpd-V3.1.1. All of workstations are configured to have fixed-addresses between .3 and .40 (the motivation behind that choice is mostly management/political much like here). DHCP leases are given out in the range of .100 to .175. Statically configured servers live in the .200 block and above (which is mostly empty). When we move to the Windows platform, management/political considerations require me to move the IP ranges around again. We would like to keep .1 - .10 reserved for network appliances, switches, and other infrastructure. .200 will remain designated for servers. The addressing space in between should be available to clients and IPs should be dynamically allocated (Edit: instead of automatic as originally mentioned) by the server. My Address Pool on the Windows Server looks like this: 192.168.0.1 192.168.0.254 (Address range for distribution) 192.168.0.1 192.168.0.10 (IP addresses excluded from distribution) 192.168.0.200 192.168.0.254 (IP addresses excluded from distribution) Currently, we have all of our clients still on the .3 - .40 range, and a few machines still active in the .100 - .175 (although there are lots devices that are powered off that still have expired leases with IPs from that range). Since the lease "database" isn't shared between the old and new DHCP server how can I prevent clients from receiving a lease with an IP address that is currently being held by client with a non-expired lease from the old DHCP server? If I just expand the range on the Debian DHCP server to be 192.168.0.10 - 192.168.0.199 is there a way to force clients to not re-use their old IP address when they send their DHCPDISCOVER? Can I make the Windows DHCP server be authoritiative like the ISC implementation? The dhcpd.conf from the Debian server: ddns-update-style none; authoritative; default-lease-time 43200; #12 hours max-lease-time 86400; #24 hours subnet 192.168.0.0 netmask 255.255.255.0 { option routers 192.168.0.1; option subnet-mask 255.255.255.0; option broadcast-address 192.168.0.255; range 192.168.0.100 192.168.0.175; } host workstation-1 { hardware ethernet 00:11:22:33:44:55; fixed-address 192.168.0.3; } ... and so on until 192.168.0.40

    Read the article

  • Windows 2003 GPO Software Restrictions

    - by joeqwerty
    We're running a Terminal Server farm in a Windows 2003 Domain, and I found a problem with the Software Restrictions GPO settings that are being applied to our TS servers. Here are the details of our configuration and the problem: All of our servers (Domain Controllers and Terminal Servers) are running Windows Server 2003 SP2 and both the domain and forest are at Windows 2003 level. Our TS servers are in an OU where we have specific GPO's linked and have inheritance blocked, so only the TS specific GPO's are applied to these TS servers. Our users are all remote and do not have workstations joined to our domain, so we don't use loopback policy processing. We take a "whitelist" approach to allowing users to run applications, so only applications that we approve and add as path or hash rules are able to run. We have the Security Level in Software Restrictions set to Disallowed and Enforcement is set to "All software files except libraries". What I've found is that if I give a user a shortcut to an application, they're able to launch the application even if it's not in the Additional Rules list of "whitelisted" applications. If I give a user a copy of the main executable for the application and they attempt to launch it, they get the expected "this program has been restricted..." message. It appears that the Software Restrictions are indeed working, except for when the user launches an application using a shortcut as opposed to launching the application from the main executable itself, which seems to contradict the purpose of using Software Restrictions. My questions are: Has anyone else seen this behavior? Can anyone else reproduce this behavior? Am I missing something in my understanding of Software Restrictions? Is it likely that I have something misconfigured in Software Restrictions? EDIT To clarify the problem a little bit: No higher level GPO's are being enforced. Running gpresults shows that in fact, only the TS level GPO's are being applied and I can indeed see my Software Restictions being applied. No path wildcards are in use. I'm testing with an application that is at "C:\Program Files\Application\executable.exe" and the application executable is not in any path or hash rule. If the user launches the main application executable directly from the application's folder, the Software Restrictions are enforced. If I give the user a shortcut that points to the application executable at "C:\Program Files\Application\executable.exe" then they are able to launch the program. EDIT Also, LNK files are listed in the Designated File Types, so they should be treated as executable, which should mean that they are bound by the same Software Restrictions settings and rules.

    Read the article

  • Building an SSL server farm

    - by dan
    I'm interested in building the the architecture in the article referenced below. I currently have a modestly-priced layer-4 load balancer and my application servers are the SSL endpoints. I want to put an SSL server farm in between my load balancer and my app servers. Then I will put another inexpensive load balancer between the SSL farm and my app servers, to do layer-7 routing. My web application has a fairly high amount of consumer traffic, that 6 servers can handle at about 50% capacity. Additionally, I have infrastructure traffic that is several orders of magnitude heavier than my consumer traffic. This is data coming in from all over the world that must integrate with my web application in real time. In total I have 18 app servers to handle all the traffic, plus 6 database servers. I will be adding 6 more app servers over the next 2 weeks and another 6 the 2 weeks after that. Conservatively, I estimate I will need to scale to 120 servers by the end of the year. My motivation right now is to separate the consumer traffic from the infrastructure traffic. The consumer traffic is higher priority than the infrastructure traffic and I cannot allow a stampede on the infrastructure side to take down my consumer-facing servers. Having a website that is always up is the top priority. However if there is a failure in one of the consumer app servers, I want to route that traffic to the servers designated for infrastructure traffic. The complication is that all the traffic is addressed using the same hostname and is nearly 100% https. The only way in my case to distinguish infrastructure from consumer traffic is by URL (poor architecture I inherited), so I need a layer 7 load balancer to be able to route. However for that to work I need either a fancy hardware-based SSL terminator or an SSL server farm as described above. Because my user base is rapidly scaling, I worry that if I go down the hardware path it will become very expensive very fast, especially since I will need 4 of everything for high availability (2 identical setups in 2 facilities). Meanwhile, the above diagram seems very flexible and more horizontally scalable. Has anyone built this before? Are there pre-built configurations? What considerations should I make and what software should I use (I've heard of people using apache with mod-ssl, nginx, and stunnel)? Also, when does it make sense to buy an expensive load balancer vs building an SSL server farm? http://1wt.eu/articles/2006_lb/index_05.html

    Read the article

< Previous Page | 15 16 17 18 19 20 21 22 23  | Next Page >