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  • Rails 3 - yield return or callback won't call in view <%= yield(:sidebar) || render('shared/sidebar'

    - by rzar
    Hey folks, I'm migrating a Website from Rails 2 (latest) to Rails 3 (beta2). Testing with Ruby 1.9.1p378 and Ruby 1.9.2dev (2010-04-05 trunk 27225) Stuck in a situation, i don't know which part will work well. Suspect yield is the problem, but don't know exactly. In my Layout Files I use the following technique quite often: app/views/layouts/application.html.erb: <%= yield(:sidebar) || render('shared/sidebar') %> For Example the partial look like: app/views/shared/_sidebar.html.erb: <p>Default sidebar Content. Bla Bla</p> Now it is time for the key part! In any view, I want to create a content_for block (optional). This can contain a pice of HTML etc. example below. If this block is set, the pice HTML inside should render in application.html.erb. If not, Rails should render the Partial at shared/_sidebar.html.erb on the right hand side. app/views/books/index.html.erb: <% content_for :sidebar do %> <strong>You have to read REWORK, a book from 37signals!</strong> <% end %> So you've got the idea. Hopefully. This technique worked well in any Rails 2.x Application. Now, in Rails 3 (beta2) only the yield Part is working. || render('shared/sidebar') The or side will not process by rails or maybe ruby. Thanks for input and time!

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  • Python: speed up removal of every n-th element from list.

    - by ChristopheD
    I'm trying to solve this programming riddle and althought the solution (see code below) works correct, it is too slow for succesful submission. Any pointers as how to make this run faster? (removal of every n-th element from a list)? Or suggestions for a better algorithm to calculate the same; seems I can't think of anything else then brute-force for now... Basically the task at hand is: GIVEN: L = [2,3,4,5,6,7,8,9,10,11,........] 1. Take the first remaining item in list L (in the general case 'n'). Move it to the 'lucky number list'. Then drop every 'n-th' item from the list. 2. Repeat 1 TASK: Calculate the n-th number from the 'lucky number list' ( 1 <= n <= 3000) My current code (it calculates the 3000 first lucky numbers in about a second on my machine - but unfortunately too slow): """ SPOJ Problem Set (classical) 1798. Assistance Required URL: http://www.spoj.pl/problems/ASSIST/ """ sieve = range(3, 33900, 2) luckynumbers = [2] while True: wanted_n = input() if wanted_n == 0: break while len(luckynumbers) < wanted_n: item = sieve[0] luckynumbers.append(item) items_to_delete = set(sieve[::item]) sieve = filter(lambda x: x not in items_to_delete, sieve) print luckynumbers[wanted_n-1]

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  • dynamic naming of UIButtons within a loop - objective-c, iphone sdk

    - by von steiner
    Dear Members, Scholars. As it may seem obvious I am not armed with Objective C knowledge. Levering on other more simple computer languages I am trying to set a dynamic name for a list of buttons generated by a simple loop (as the following code suggest). Simply putting it, I would like to have several UIButtons generated dynamically (within a loop) naming them dynamically, as well as other related functions. button1,button2,button3 etc.. After googling and searching Stackoverlow, I haven't arrived to a clear simple answer, thus my question. - (void)viewDidLoad { // This is not Dynamic, Obviously UIButton *button0 = [UIButton buttonWithType:UIButtonTypeRoundedRect]; [button0 setTitle:@"Button0" forState:UIControlStateNormal]; button0.tag = 0; button0.frame = CGRectMake(0.0, 0.0, 100.0, 100.0); button0.center = CGPointMake(160.0,50.0); [self.view addSubview:button0]; // I can duplication the lines manually in terms of copy them over and over, changing the name and other related functions, but it seems wrong. (I actually know its bad Karma) // The question at hand: // I would like to generate that within a loop // (The following code is wrong) float startPointY = 150.0; // for (int buttonsLoop = 1;buttonsLoop < 11;buttonsLoop++){ NSString *tempButtonName = [NSString stringWithFormat:@"button%i",buttonsLoop]; UIButton tempButtonName = [UIButton buttonWithType:UIButtonTypeRoundedRect]; [tempButtonName setTitle:tempButtonName forState:UIControlStateNormal]; tempButtonName.tag = tempButtonName; tempButtonName.frame = CGRectMake(0.0, 0.0, 100.0, 100.0); tempButtonName.center = CGPointMake(160.0,50.0+startPointY); [self.view addSubview:tempButtonName]; startPointY += 100; } }

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  • Best way to convert wpf triggers to silverlight?

    - by Stewart Armbrecht
    I have developed several custom controls in a wpf application that use triggers. what is the fastest way to convert the code so that I have a single code base that works both in the wpf application and the silverlight applicaiton. Here is a sample of the code: <Style x:Key="sButtonAction" TargetType="Button"> <!--<Setter Property="BitmapEffect" Value="{StaticResource BannerEffect}" />--> <Setter Property="Height" Value="25" /> <Setter Property="Margin" Value="4" /> <Setter Property="Cursor" Value="Hand" /> <Setter Property="Template"> <Setter.Value> <ControlTemplate TargetType="Button"> <Border x:Name="PART_Border" CornerRadius="10" BorderThickness="{StaticResource sBorderThicknessStandard}" BorderBrush="{StaticResource bColorBorder}" Background="{StaticResource ButtonActionBackground}"> <TextBlock x:Name="PART_TextBlock" Margin="5,2,5,2" HorizontalAlignment="Center" VerticalAlignment="Center" Foreground="White"> <ContentPresenter HorizontalAlignment="Center" VerticalAlignment="Center" /></TextBlock> </Border> <ControlTemplate.Triggers> <Trigger Property="IsMouseOver" Value="True"> <Setter TargetName="PART_TextBlock" Property="Foreground" Value="#990000"></Setter> </Trigger> <Trigger Property="IsPressed" Value="True"> <Setter TargetName="PART_Border" Property="Background" Value="{StaticResource ButtonActionBackgroundSelected}"></Setter> </Trigger> </ControlTemplate.Triggers> </ControlTemplate> </Setter.Value> </Setter> </Style>

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  • Authlogic and password and password confirmation attributes - inaccessible?

    - by adam
    Im trying to test my successfully creates a new user after login (using authlogic). Ive added a couple of new fields to the user so just want to make sure that the user is saved properly. The problem is despite creating a valid user factory, whenever i try to grab its attributes to post to the create method, password and password confirmation are being ommitted. I presuem this is a security method that authlogic performs in the background. This results in validations failing and the test failing. Im wondering how do i get round this problem? I could just type the attributes out by hand but that doesnt seem very dry. context "on POST to :create" do context "on posting a valid user" do setup do @user = Factory.build(:user) post :create, :user => @user.attributes end should "be valid" do assert @user.valid? end should_redirect_to("users sentences index page") { sentences_path() } should "add user to the db" do assert User.find_by_username(@user.username) end end ##User factory Factory.define :user do |f| f.username {Factory.next(:username) } f.email { Factory.next(:email)} f.password_confirmation "password" f.password "password" f.native_language {|nl| nl.association(:language)} f.second_language {|nl| nl.association(:language)} end

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  • Getting started with SVG graphics objects in JSF 2.0 pages.

    - by AlanObject
    What I want to do is create web pages with interactive SVG content. I had this working as a Java desktop application using Batik to render my SVG and collect UI events like mouseclick. Now I want to use those SVG graphics files in my JSF (Primefaces) web application in the same way. Trying to get started, I found this didn't work: <h:graphicImage id="gloob" value="images/sprinkverks.svg" alt="Graphic Goes Here"/> I don't mind doing some reading to get up the learning curve. It was just a bit surprising that some google searches didn't turn up anything useful. What I did find suggested that I would have to do this with the f:verbatim tag as if I were hand-coding the HTML. I would then have to add some script to capture the SVG events and feed them back into the AJAX code. If I have to do all that I will, but I was hoping there would be an easier and automated way. So the questions are: How to get the image to render in the first place? How to get the DOM events from the SVG portion of the page back to the backing beans? Much thanks for any pointers.

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  • NSFetchedResultsController not updating UITableView's section indexes

    - by Luther Baker
    I am populating a UITableViewController with an NSFetchedResultsController with results creating sections that populate section headers and a section index. I am using the following method to populate the section index: - (NSArray *)sectionIndexTitlesForTableView:(UITableView *)tableView { return [fetchedResultsController_ sectionIndexTitles]; } and now I've run into a problem. When I add a new element to the NSManagedObjectContext associated with the NSFetchedResultsController, the new element is saved and appropriately displayed as a cell in the UITableView ... except for one thing. If the new element creates a new SECTION, the new section index does not show up in the right hand margin unless I pop the UINavigationController's stack and reload the UITableViewController. I have conformed to the NSFetchedResultsControllerDelegate's interface and manually invoke [self.tableView reloadSectionIndexTitles]; at the end of both these delegate methods: controller:didChangeSection... controller:didChangeObject... and while I can debug and trace the execution into the methods and see the reload call invoked, the UITableView's section index never reflects the section changes. Again, the data shows up - new sections are physically visible (or removed) in the UITableView but the section indexes are not updated. Am I missing something?

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  • Attempting to find a formula for tessellating rectangles onto a board, where middle square can't be

    - by timemirror
    I'm working on a spatial stacking problem... at the moment I'm trying to solve in 2D but will eventually have to make this work in 3D. I divide up space into n x n squares around a central block, therefore n is always odd... and I'm trying to find the number of locations that a rectangle of any dimension less than n x n (eg 1x1, 1x2, 2x2 etc) can be placed, where the middle square is not available. So far I've got this.. total number of rectangles = ((n^2 + n)^2 ) / 4 ..also the total number of squares = (n (n+1) (2n+1)) / 6 However I'm stuck in working out a formula to find how many of those locations are impossible as the middle square would be occupied. So for example: [] [] [] [] [x] [] [] [] [] 3 x 3 board... with 8 possible locations for storing stuff as mid square is in use. I can use 1x1 shapes, 1x2 shapes, 2x1, 3x1, etc... Formula gives me the number of rectangles as: (9+3)^2 / 4 = 144/4 = 36 stacking locations However as the middle square is unoccupiable these can not all be realized. By hand I can see that these are impossible options: 1x1 shapes = 1 impossible (mid square) 2x1 shapes = 4 impossible (anything which uses mid square) 3x1 = 2 impossible 2x2 = 4 impossible etc Total impossible combinations = 16 Therefore the solution I'm after is 36-16 = 20 possible rectangular stacking locations on a 3x3 board. I've coded this in C# to solve it through trial and error, but I'm really after a formula as I want to solve for massive values of n, and also to eventually make this 3D. Can anyone point me to any formulas for these kind of spatial / tessellation problem? Also any idea on how to take the total rectangle formula into 3D very welcome! Thanks!

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  • Source Control Manager Backend

    - by Gabriel Parenza
    Hi Friends, What do you think is a better approach for Source Control Manager Backend. I am weighing File system vs Hosted Subversion service. Hosted Subversion-- (My company already has another group taking care of this) Advantages: * Zero maintenance on our end * Auto-backup and recovery * Reliability by auto-backup and file redundancy. * File history view in built, file merge, file diff On the other hand, while File system does not have the featured mentioned above but is much more simpler. Moreover, if files are hosted on Linux machine, which is backed up, it takes care of file system crash issues. Subversion will need working copies, which are going to be on this same Linux machine, and hence the need to not have an extra layer. Folks, I am looking for stronger reasons why I should take Subversion instead of keeping thing simple and going with File System. Let me know your opinions. Very thanks in advance, Gabriel. PS: I have explored few Commercial Source Manager, and have decide to go this route as it better suits our need.

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  • Nested dereferencing arrows in Perl: to omit or not to omit?

    - by DVK
    In Perl, when you have a nested data structure, it is permissible to omit de-referencing arrows to 2d and more level of nesting. In other words, the following two syntaxes are identical: my $hash_ref = { 1 => [ 11, 12, 13 ], 3 => [31, 32] }; my $elem1 = $hash_ref->{1}->[1]; my $elem2 = $hash_ref->{1}[1]; # exactly the same as above Now, my question is, is there a good reason to choose one style over the other? It seems to be a popular bone of stylistic contention (Just on SO, I accidentally bumped into this and this in the space of 5 minutes). So far, none of the usual suspects says anything definitive: perldoc merely says "you are free to omit the pointer dereferencing arrow". Conway's "Perl Best Practices" says "whenever possible, dereference with arrows", but it appears to only apply to the context of dereferencing the main reference, not optional arrows on 2d level of nested data structures. "MAstering Perl for Bioinfirmatics" author James Tisdall doesn't give very solid preference either: "The sharp-witted reader may have noticed that we seem to be omitting arrow operators between array subscripts. (After all, these are anonymous arrays of anonymous arrays of anonymous arrays, etc., so shouldn't they be written [$array-[$i]-[$j]-[$k]?) Perl allows this; only the arrow operator between the variable name and the first array subscript is required. It make things easier on the eyes and helps avoid carpal tunnel syndrome. On the other hand, you may prefer to keep the dereferencing arrows in place, to make it clear you are dealing with references. Your choice." Personally, i'm on the side of "always put arrows in, since itg's more readable and obvious tiy're dealing with a reference".

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  • Front End Developer v/s PHP-MySQL Engineer

    - by user301943
    Hello, I want to decide which of this would be a more viable career option? I am ready to quit my current job and hence I am looking for new opportunity. Current job is maintainence and no more active development. My current role is of a PHP/MySQL Developer. I very well understand web-programming and am comfortable with RoR/Sinatra/Zend MVC/JQuery/JSON manipulation, etc. I understand MySQL InnoDB/MyISAM engine and how one differs from the other, etc. Basically, I could very well manage the deployment of a web-application end-to-end including configuration of Apache/Nginx servers, memcache,etc On the other hand, I am being offered a Sr.Front End Web developer that would require me to extensively write HTML/CSS crossbrowser/crossplatform compliant code. I very well understand XHTML/CSS/Box model etc. I would be working on Drupal for the management of websites. While I understand continuing to work on server-side technologies would always be a good career path, how would the role of Core front-end developer turn out to be? If I take this opportunity, will I eventually get a chance to focus onto UCD, HCI, Information Architect,etc. So are these kinda roles possible if I focus on front end development? No offenses to the Front end developers, just want to understand if this is something I want to gain a mastery over. I have 2 yrs of industry experience after graduating with a MS-Computer Science. Although, I have a CS degree, if I were to take uip serious front-end role; I could probably go back and take up some design/HCI/UI courses. Please advise.

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  • Javascript :(….. Oh!! So its jquery? Now what?? I’m a C# Guy

    - by Shekhar_Pro
    Hi guys I want you to Guide me here. This other day I was working out some AJAX for my ASP.Net website and handling client side code in Java was taking the hell out of me. Then I got my hands on this Book jQuery In Action 2nd Edition and solved my problem with the help of Example code in the book. Now as I checked the contents I got an overview that whatever I had ever thought of doing can be done by this jQuery so easily and quite cleanly. I am actually pretty new to web development (say abt 4months ) and from C# world where we have cool libraries and Simple and Elegant coding style. (yeah including those generic, Ienumerable, lambadas, chained statements.. you got it…) and you know what you’re doing when writing some code. And we have so great IntelliSense to care., and above all we have everything Strongly Typed. But in Javascript everything is so messy.. . (and I don’t know why they are not properly indented.. see page source ) Now tell me what should I do, go straight with jQuery or should I first learn Javascript (like a disciplined boy…I even have a book for that too… got in gift :) …. ) I have seen Is it a good idea to learn JavaScript before learning jQuery? but remember I have already got a project on my hand…

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  • SQL 2005 indexed queries slower than unindexed queries

    - by uos??
    Adding a seemingly perfectly index is having an unexpectedly adverse affect on a query performance... -- [Data] has a predictable structure and a simple clustered index of the primary key: ALTER TABLE [dbo].[Data] ADD PRIMARY KEY CLUSTERED ( [ID] ) -- My query, joins on itself looking for a certain kind of "overlapping" records SELECT DISTINCT [Data].ID AS [ID] FROM dbo.[Data] AS [Data] JOIN dbo.[Data] AS [Compared] ON [Data].[A] = [Compared].[A] AND [Data].[B] = [Compared].[B] AND [Data].[C] = [Compared].[C] AND ([Data].[D] = [Compared].[D] OR [Data].[E] = [Compared].[E]) AND [Data].[F] <> [Compared].[F] WHERE 1=1 AND [Data].[A] = @A AND @CS <= [Data].[C] AND [Data].[C] < @CE -- Between a range [Data] has about a quarter-million records so far, 10% to 50% of the data satisfies the where clause depending on @A, @CS, and @CE. As is, the query takes 1 second to return about 300 rows when querying 10%, and 30 seconds to return 3000 rows when querying 50% of the data. Curiously, the estimated/actual execution plan indicates two parallel Clustered Index Scans, but the clustered index is only of the ID, which isn't part of the conditions of the query, only the output. ?? If I add this hand-crafted [IDX_A_B_C_D_E_F] index which I fully expected to improve performance, the query slows down by a factor of 8 (8 seconds for 10% & 4 minutes for 50%). The estimated/actual execution plans show an Index Seek, which seems like the right thing to be doing, but why so slow?? CREATE UNIQUE INDEX [IDX_A_B_C_D_E_F] ON [dbo].[Data] ([A], [B], [C], [D], [E], [F]) INCLUDE ([ID], [X], [Y], [Z]); The Data Engine Tuning wizard suggests a similar index with no noticeable difference in performance from this one. Moving AND [Data].[F] <> [Compared].[F] from the join condition to the where clause makes no difference in performance. I need these and other indexes for other queries. I'm sure I could hint that the query should refer to the Clustered Index, since that's currently winning - but we all know it is not as optimized as it could be, and without a proper index, I can expect the performance will get much worse with additional data. What gives?

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  • What makes static initialization functions good, bad, or otherwise?

    - by Richard Levasseur
    Suppose you had code like this: _READERS = None _WRITERS = None def Init(num_readers, reader_params, num_writers, writer_params, ...args...): ...logic... _READERS = new ReaderPool(num_readers, reader_params) _WRITERS = new WriterPool(num_writers, writer_params) ...more logic... class Doer: def __init__(...args...): ... def Read(self, ...args...): c = _READERS.get() try: ...work with conn finally: _READERS.put(c) def Writer(...): ...similar to Read()... To me, this is a bad pattern to follow, some cons: Doers can be created without its preconditions being satisfied The code isn't easily testable because ConnPool can't be directly mocked out. Init has to be called right the first time. If its changed so it can be called multiple times, extra logic has to be added to check if variables are already defined, and lots of NULL values have to be passed around to skip re-initializing. In the event of threads, the above becomes more complicated by adding locking Globals aren't being used to communicate state (which isn't strictly bad, but a code smell) On the other hand, some pros: its very convenient to call Init(5, "user/pass", 2, "user/pass") It simple and "clean" Personally, I think the cons outweigh the pros, that is, testability and assured preconditions outweigh simplicity and convenience.

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  • Binding a member signal to a function

    - by the_drow
    This line of code compiles correctly without a problem: boost::bind(boost::ref(connected_), boost::dynamic_pointer_cast<session<version> >(shared_from_this()), boost::asio::placeholders::error); However when assigning it to a boost::function or as a callback like this: socket_->async_connect(connection_->remote_endpoint(), boost::bind(boost::ref(connected_), boost::dynamic_pointer_cast<session<version> >(shared_from_this()), boost::asio::placeholders::error)); I'm getting a whole bunch of incomprehensible errors (linked since it's too long to fit here). On the other hand I have succeeded binding a free signal to a boost::function like this: void print(const boost::system::error_code& error) { cout << "session connected"; } int main() { boost::signal<void(const boost::system::error_code &)> connected_; connected_.connect(boost::bind(&print, boost::asio::placeholders::error)); client<>::connection_t::socket_ptr socket_(new client<>::connection_t::socket_t(conn->service())); // shared_ptr of a tcp socket socket_->async_connect(conn->remote_endpoint(), boost::bind(boost::ref(connected_), boost::asio::placeholders::error)); conn->service().run(); // io_service.run() return 0; } This works and prints session connected correctly. What am I doing wrong here?

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  • What scalability problems have you solved using a NoSQL data store?

    - by knorv
    NoSQL refers to non-relational data stores that break with the history of relational databases and ACID guarantees. Popular open source NoSQL data stores include: Cassandra (tabular, written in Java, used by Facebook, Twitter, Digg, Rackspace, Mahalo and Reddit) CouchDB (document, written in Erlang, used by Engine Yard and BBC) Dynomite (key-value, written in C++, used by Powerset) HBase (key-value, written in Java, used by Bing) Hypertable (tabular, written in C++, used by Baidu) Kai (key-value, written in Erlang) MemcacheDB (key-value, written in C, used by Reddit) MongoDB (document, written in C++, used by Sourceforge, Github, Electronic Arts and NY Times) Neo4j (graph, written in Java, used by Swedish Universities) Project Voldemort (key-value, written in Java, used by LinkedIn) Redis (key-value, written in C, used by Engine Yard, Github and Craigslist) Riak (key-value, written in Erlang, used by Comcast and Mochi Media) Ringo (key-value, written in Erlang, used by Nokia) Scalaris (key-value, written in Erlang, used by OnScale) ThruDB (document, written in C++, used by JunkDepot.com) Tokyo Cabinet/Tokyo Tyrant (key-value, written in C, used by Mixi.jp (Japanese social networking site)) I'd like to know about specific problems you - the SO reader - have solved using data stores and what NoSQL data store you used. Questions: What scalability problems have you used NoSQL data stores to solve? What NoSQL data store did you use? What database did you use before switching to a NoSQL data store? I'm looking for first-hand experiences, so please do not answer unless you have that.

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  • rails + compass: advantages vs using haml + blueprint directly

    - by egarcia
    I've got some experience using haml (+sass) on rails projects. I recently started using them with blueprintcss - the only thing I did was transform blueprint.css into a sass file, and started coding from there. I even have a rails generator that includes all this by default. It seems that Compass does what I do, and other things. I'm trying to understand what those other things are - but the documentation/tutorials weren't very clear. These are my conclusions: Compass comes with built-in sass mixins that implement common CSS idioms, such as links with icons or horizontal lists. My solution doesn't provide anything like that. (1 point for Compass). Compass has several command-line options: you can create a rails project, but you can also "install" it on an existing rails project. A rails generator could be personalized to do the same thing, I guess. (Tie). Compass has two modes of working with blueprint: "basic" and "semantic" usage. I'm not clear about the differences between those. With my rails generator I only have one mode, but it seems enough. (Tie) Apparently, Compass is prepared to use other frameworks, besides blueprint (e.g. YUI). I could not find much documentation about this, and I'm not interested on it anyway - blueprint is ok for me (Tie). Compass' learning curve seems a bit stiff and the documentation seems sparse. Learning could be a bit difficult. On the other hand, I know the ins and outs of my own system and can use it right away. (1 point for my system). With this analysis, I'm hesitant to give Compass a try. Is my analysis correct? Are Am I missing any key points, or have I evaluated any of these points wrongly?

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  • toggling proximity sensor on iPhone loses an event

    - by slugolicious
    I'm using setProximitySensingEnabled and implemented proximityStateChanged in my UIApplication subclass. It looks like if sensing is toggled, that the first "off" event is being lost. My UIApplication class is pretty basic... -(void)proximityStateChanged:(BOOL)state { NSLog(state ? @"ON" : @"OFF"); } In my application delegate, I have a UISwitch that enables/disables the proximity sensor. -(IBAction)toggleProxy:(id)sender { [UIApplication sharedApplication].proximitySensingEnabled = prox.on; } "prox" is my UISwitch. The test works fine when it first starts. I tap the switch to turn it on and then put my hand over the sensor for a second then move it away and get: 2009-03-11 12:43:00.465 Proximity[324:20b] ON 2009-03-11 12:43:02.514 Proximity[324:20b] OFF 2009-03-11 12:43:04.046 Proximity[324:20b] ON 2009-03-11 12:43:05.621 Proximity[324:20b] OFF I then tap the switch to turn it off then tap again to turn it on. Now I get: 2009-03-11 12:43:12.005 Proximity[324:20b] ON 2009-03-11 12:43:14.789 Proximity[324:20b] ON 2009-03-11 12:43:16.467 Proximity[324:20b] OFF 2009-03-11 12:43:17.516 Proximity[324:20b] ON 2009-03-11 12:43:19.077 Proximity[324:20b] OFF Notice I get two ON's before an OFF. The OFF is lost somewhere. I can't replicate this behavior using Google's mobile app so I'm wondering if they're resetting something in between proximity enabling. They don't have the proximity sensor on all the time because if you cover the sensor, the screen doesn't go blank. You have to tilt the phone up and angle it back (to simulate the position it would be in at your ear) and then covering the sensor works. Anyone else playing with the sensor? In my particular app, I'm recording a voice message and when you move the phone away from your ear, I want to pause the recording (when I get an OFF). The first time I move the phone away from my ear, the recording is not paused. However, if I put it to my ear and move it away again, it is paused.

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  • Table design issues - should I create separate fields or store as a blob

    - by Ali
    Hi guys I'm working on my web based ordering system and we would like to maintain a kind of task history for each of our orders. A hsitory in the sense that we would like to maintain a log of who did what on an order like lets say an order has been entered - we would like to know if the order was acknowledged for an example. Or lets say somebody followed up on the order - etc. Consider that there are numerous situations like this for each order would it be wise to create a schema on the lines of: Orders ID - title - description - date - is_ack - is_follow - ack_by ..... That accounts to a lot of fields - on teh other hand I could have one LongText field called 'history' and fill it with a serialised object holding all the information. However in the latter case I can't run a query to lets say retrieve all orders that have not been acknowledged and stuff like that. With time requirements woudl change and I would be required to modify it to allow for more detailed tracking and that is why I need to set up a way which would be feasible to scale upon yet I don't want to be restricted on the SQL side too much.

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  • 'Must Override a Superclass Method' Errors after importing a project into Eclipse

    - by Tim H
    Anytime I have to re-import my projects into Eclipse (if I reinstalled Eclipse, or changed the location of the projects), almost all of my overridden methods are not formatted correctly, causing the error 'The method ?????????? must override a superclass method'. It may be noteworthy to mention this is with Android projects - for whatever reason, the method argument values are not always populated, so I have to manually populate them myself. For instance: list.setOnCreateContextMenuListener(new OnCreateContextMenuListener() { public void onCreateContextMenu(ContextMenu menu, View v, ContextMenuInfo menuInfo) { //These arguments have their correct names } }); will be initially populated like this: list.setOnCreateContextMenuListener(new OnCreateContextMenuListener() { public void onCreateContextMenu(ContextMenu arg1, View arg2, ContextMenuInfo arg3) { //This methods arguments were not automatically provided } }); The odd thing is, if I remove my code, and have Eclipse automatically recreate the method, it uses the same argument names I already had, so I don't really know where the problem is, other then it auto-formatting the method for me. This becomes quite a pain having to manually recreate ALL my overridden methods by hand. If anyone can explain why this happens or how to fix it .. I would be very happy. Maybe it is due to the way I am formatting the methods, which are inside an argument of another method?

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  • To (monkey)patch or not to (monkey)patch, that is the question

    - by gsakkis
    I was talking to a colleague about one rather unexpected/undesired behavior of some package we use. Although there is an easy fix (or at least workaround) on our end without any apparent side effect, he strongly suggested extending the relevant code by hard patching it and posting the patch upstream, hopefully to be accepted at some point in the future. In fact we maintain patches against specific versions of several packages that are applied automatically on each new build. The main argument is that this is the right thing to do, as opposed to an "ugly" workaround or a fragile monkey patch. On the other hand, I favor practicality over purity and my general rule of thumb is that "no patch" "monkey patch" "hard patch", at least for anything other than a (critical) bug fix. So I'm wondering if there is a consensus on when it's better to (hard) patch, monkey patch or just try to work around a third party package that doesn't do exactly what one would like. Does it have mainly to do with the reason for the patch (e.g. fixing a bug, modifying behavior, adding missing feature), the given package (size, complexity, maturity, developer responsiveness), something else or there are no general rules and one should decide on a case-by-case basis ?

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  • In C# should I reuse a function / property parameter to compute temp result or create a temporary v

    - by Hamish Grubijan
    The example below may not be problematic as is, but it should be enough to illustrate a point. Imagine that there is a lot more work than trimming going on. public string Thingy { set { // I guess we can throw a null reference exception here on null. value = value.Trim(); // Well, imagine that there is so much processing to do this.thingy = value; // That this.thingy = value.Trim() would not fit on one line ... So, if the assignment has to take two lines, then I either have to abusereuse the parameter, or create a temporary variable. I am not a big fan of temporary variables. On the other hand, I am not a fan of convoluted code. I did not include an example where a function is involved, but I am sure you can imagine it. One concern I have is if a function accepted a string and the parameter was "abused", and then someone changed the signature to ref in both places - this ought to mess things up, but ... who would knowingly make such a change if it already worked without a ref? Seems like it is their responsibility in this case. If I mess with the value of value, am I doing something non-trivial under the hood? If you think that both approaches are acceptable, then which do you prefer and why? Thanks.

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  • What is the fastest way to pull a few element values out of XML files in Perl?

    - by Anon Guy
    I have a bunch of XML files that are about 1-2 megabytes in size. Actually, more than a bunch, there are millions. They're all well-formed and many are even validated against their schema (confirmed with libxml2). All were created by the same app, so they're in a consistent format (though this could theoretically change in the future). I want to check the values of one element in each file from within a Perl script. Speed is important (I'd like to take less than a second per file) and as noted I already know the files are well-formed. I am sorely tempted to simply 'open' the files in Perl and scan through until I see the element I am looking for, grab the value (which is near the start of the file), and close the file. On the other hand, I could use an XML parser (which might protect me from future changes to the XML formatting) but I suspect it will be slower than I'd like. Can anyone recommend an appropriate approach and/or parser? Thanks in advance.

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  • Correct use of WSDL-generated sources

    - by John K
    How can I easily convert between manually written classes and WSDL-generated equivalents? I have a Java SE 6 thick client that calls a web service to get and store data. The client has a DAO that works with my entity classes, calls <Entity.toDto() to convert them to DTOs, and sends/receives that data with the web service. My issue stems from the fact that the entity classes live on both sides of the service interface: client and server. Each entity has a constructor from the DTO and a toDto function: public class EntityClass { public EntityClass(EntityClassDto dto); public EntityClassDto toDto(); ... } This means I have a handwritten DTO class that the client and server both use. However, the service interface expects the WSDL-generated classes. I have tried writing conversion code between the hand-written DTO and the WSDL-generated DTO and it is tedious and error-prone. What is a reasonable alternative to this? Some back-story: The thick client should be able to have a configurable backend: either direct to the DB or through this web service. The aforementioned DAO is the web service based implementation and another imlpementation that is JPA-based exists.

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  • How to estimate the contribution of an individual to a software project?

    - by Amit Kumar
    I work on a software project and would like to estimate the percentage out of the total contribution that I have put in the development of the software. Is there some tool doing this? Such a tool can be useful for appraisals or negotiations, for example. After all, we work for money (yes, not only money, put the point remains). I think there is enough hand-waving for the most important things. The estimation is very subjective (at least to me now) but I do not know of any tool that provides even a subjective estimate. I know of Sloccount that spells out the total effort using the lines of code but not on per-developer basis. My idea of an ideal tool for this purpose would: measure the complexity of the code (more complex is more effort, but more effort is not necessarily more contribution) measure the decomposibility/flexibility of the software (more decomposable is better) how much library code is used -- using library code speeds up the development process, increases the associated risk and requires the developer to know from before or learn about the library. be intelligent enough to differentiate between "who wrote the code", "who copied the code" and "who indented the code". It is difficult to differentiate between the complexity in the implementation and the intrinsic complexity of the problem. Perhaps a comparison can be made with an equivalent open source counterpart if there is, or for each submodule separately. If there is no such tool, is there no merit in having such a tool? Or do you believe in "I do work, I do not measure"? It takes time after all. Perhaps the project manager should do this estimation continuously, say, weekly. Are there any standards? Yes, standardization is difficult because every project has a different goal, but difficult does not mean it is not useful.

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