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  • Why is this statement treated as a string instead of its result?

    - by reve_etrange
    I am trying to perform some composition-based filtering on a large collection of strings (protein sequences). I wrote a group of three subroutines in order to take care of it, but I'm running into trouble in two ways - one minor, one major. The minor trouble is that when I use List::MoreUtils 'pairwise' I get warnings about using $a and $b only once and them being uninitialized. But I believe I'm calling this method properly (based on CPAN's entry for it and some examples from the web). The major trouble is an error "Can't use string ("17/32") as HASH ref while "strict refs" in use..." It seems like this can only happen if the foreach loop in &comp is giving the hash values as a string instead of evaluating the division operation. I'm sure I've made a rookie mistake, but can't find the answer on the web. The first time I even looked at perl code was last Wednesday... use List::Util; use List::MoreUtils; my @alphabet = ( 'A', 'R', 'N', 'D', 'C', 'Q', 'E', 'G', 'H', 'I', 'L', 'K', 'M', 'F', 'P', 'S', 'T', 'W', 'Y', 'V' ); my $gapchr = '-'; # Takes a sequence and returns letter = occurrence count pairs as hash. sub getcounts { my %counts = (); foreach my $chr (@alphabet) { $counts{$chr} = ( $[0] =~ tr/$chr/$chr/ ); } $counts{'gap'} = ( $[0] =~ tr/$gapchr/$gapchr/ ); return %counts; } # Takes a sequence and returns letter = fractional composition pairs as a hash. sub comp { my %comp = getcounts( $[0] ); foreach my $chr (@alphabet) { $comp{$chr} = $comp{$chr} / ( length( $[0] ) - $comp{'gap'} ); } return %comp; } # Takes two sequences and returns a measure of the composition difference between them, as a scalar. # Originally all on one line but it was unreadable. sub dcomp { my @dcomp = pairwise { $a - $b } @{ values( %{ comp( $[0] ) } ) }, @{ values( %{ comp( $[1] ) } ) }; @dcomp = apply { $_ ** 2 } @dcomp; my $dcomp = sqrt( sum( 0, @dcomp ) ) / 20; return $dcomp; } Much appreciation for any answers or advice!

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  • MS Access 07 - Q re lookup column vs many-to-many; Q re checkboxes in many-to-many forms

    - by TBinLondon
    Hello, I'm creating a database with Access. This is just a test database, similar to my requirements, so I can get my skills up before creating one for work. I've created a database for a fictional school as this is a good playground and rich data (many students have many subjects have many teachers, etc). Question 1 What is the difference, if any, between using a Lookup column and a many-to-many associate table? Example: I have Tables 'Teacher' and 'Subject'. Many teachers have many subjects. I can, and have, created a table 'Teacher_Subject' and run queries with this. I have then created a lookup column in teachers table with data from subjects. The lookup column seems to take the place of the teacher_subject table. (though the data on relationships is obviously duplicated between lookup table and teacher_subject and may vary). Which one is the 'better' option? Is there a snag with using lookup tables? (I realize that this is a very 'general' question. Links to other resources and answers saying 'that depends...' are appreciated) Question 2 What attracts me to lookup tables is the following: When creating a form for entering subjects for teachers, with lookup I can simply create checkboxes and click a subject for a teacher 'on' or 'off'. Each click on/off creates/removes a record in the lookup column (which replaces teacher_subject). If I use a form from a query from teacher subject with teacher as main form and subject as subform I run into this problem: In the subform I can either select each subject that teacher has in a bombo box, i.e. click, scroll down, select, go to next row, click, scroll down, etc. (takes too long) OR I can create a list box listing all available subjects in each row but allowing me to select only one. (takes up too much space). Is it possible to have a click on/off list box for teacher_subject, creating/removing a record there with each click? Note - I know zero SQL or VB. If the correct answer is "you need to know SQL for this" then that's cool. I just need to know. Thanks!

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  • Which network protocol to use for lightweight notification of remote apps (Delphi 2005)

    - by Chris Thornton
    I have this situation.... Client-initiated SOAP 1.1 communication between one server and let's say, tens of thousands of clients. Clients are external, coming in through our firewall, authenticated by certificate, https, etc.. They can be anywhere, and usually have their own firewalls, NAT routers, etc... They're truely external, not just remote corporate offices. They could be in a corporate/campus network, DSL/Cable, even Dialup. Currently, clients push new data to the server and pull new data from the server on 15-minute polling loop. The server currently does not push data - the client hits the "messagecount" method, to see if there is new data to pull. If 0, it sleeps for another 15 min and checks again. We're trying to get that down to 7 seconds. If this were an internal app, with one or just a few dozen clients, we'd write a cilent "listener" soap service, and would push data to it. But since they're external, sit behind their own firewalls, and sometimes private networks behind NAT routers, this is not practical. So we're left with polling on a much quicker loop. 10K clients, each checking their messagecount every 10 seconds, is going to be 1000/sec messages that will mostly just waste bandwidth, server, firewall, and authenticator resources. So I'm trying to design something better than what would amount to a self-inflicted DoS attack. I don't think it's practical to have the server send soap messages to the client (push) as this would require too much configuration at the client end. But I think there are alternatives that I don't know about. Such as: 1) Is there a way for the client to make a request for GetMessageCount() via Soap 1.1, and get the response, and then perhaps, "stay on the line" for perhaps 5-10 minutes to get additional responses in case new data arrives? i.e the server says "0", then a minute later in response to some SQL trigger (the server is C# on Sql Server, btw), knows that this client is still "on the line" and sends the updated message count of "5"? 2) Is there some other protocol that we could use to "ping" the client, using information gathered from their last GetMessageCount() request? 3) I don't even know. I guess I'm looking for some magic protocol where the client can send a GetMessageCount() request, which would include info for "oh by the way, in case the answer changes in the next hour, ping me at this address...". Also, I'm assuming that any of these "keep the line open" schemes would seriously impact the server sizing, as it would need to keep many thousands of connections open, simultaneously. That would likely impact the firewalls too, I think. Is there anything out there like that? Or am I pretty much stuck with polling? TIA, Chris

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  • Java - Layering issues with Lists and Graphics2D

    - by Mirrorcrazy
    So I have a DisplayPanel class that extends JPanel and that also paints my numerous images for my program using Graphics2D. In order to be able to easily customly use this I set it up so that every time the panel is repainted it uses a List, that I can add to or remove from as the program processes. My problem is with layering. I've run into an issue where the List must have reached its resizing point (or something whacky like that) and so the images i want to display end up beneath all of the other images already on the screen. I've come to the community for an answer because I have faith you will provide a good one. DisplayPanel: package earthworm; import java.awt.Graphics; import java.awt.Graphics2D; import java.util.ArrayList; import java.util.List; import javax.swing.JPanel; public class DisplayPanel extends JPanel { private List<ImageMap> images = new ArrayList(); public DisplayPanel() { setSize(800, 640); refresh(); } public void refresh() { revalidate(); repaint(); } @Override public void paintComponent(Graphics g) { super.paintComponent(g); Graphics2D g2d = (Graphics2D) g; g2d.clearRect(0, 0, 800, 640); for(int i = 0; i < images.size(); i++) g2d.drawImage( images.get(i).getImage(), images.get(i).getX(), images.get(i).getY(), null); } public void paintImage(ImageMap[] images, ImageMap[] clearImages, boolean clear) { if(clear) this.images.clear(); else if(clearImages!=null) for(int i = 0; i < clearImages.length; i++) this.images.remove(clearImages[i]); if(images!=null) for(int i = 0; i<images.length; i++) this.images.add(images[i]); refresh(); } }

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  • current_user.user_type_id = @employer ID

    - by sscirrus
    I am building a system with a User model (authenticated using AuthLogic) and three user types in three models: one of these models is Employer. Each of these three models has_many :users, :as = :authenticable. I start by having a new visitor to the site create their own 'User' record with username, password, which user type they are, etc. Upon creation, the user is sent to the 'new' action for one of the three models. So, if they tell us they are an employer, we redirect_to :controller = "employers, :action = "new". Question: When the employer has submitted, I want to set the current_user.user_type_id equal to the employer ID. This should be simple... but it's not working. # Employers Controller / new def new @employer = Employer.new 1.times {@employer.addresses.build} render :layout => 'forms' end # Employers Controller / create def create @employer = Employer.new(params[:employer]) if @employer.save if current_user.blank? redirect_to :controller => "users", :action => "new" else current_user.user_type_id = @employer.id current_user.user_type = "Employer" redirect_to :action => "home", :id => current_user.user_type_id end else render :action => "new" end end ------UPDATE------ Hi guys. In response: I am using this table structure because each of my three user type models have lots of different fields and each has different relationships to the other models, which is why I've avoided STI. By 1.times (@employer.addresses.build) I'm connecting the employer model to the address polymorphic table in one form, so I'm asking the controller to build a new address to go along with the new employer. Averell: you mentioned encapsulating... something in the model using a 'setter' method. I have no idea what you mean by this - could you please explain how this works (or direct me to an example elsewhere)? With tsdbrown's answer I have managed to create the behavior I want... if there's a more elegant way to accomplish the same thing I'd love to learn how. Thanks very much. Thanks to tsdbrown for answering the current_user.save problem!

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  • c++ recursion with arrays

    - by Sam
    I have this project that I'm working on for a class, and I'm not looking for the answer, but maybe just a few tips since I don't feel like I'm using true recursion. The problem involves solving the game of Hi Q or "peg solitaire" where you want the end result to be one peg remaining (it's played with a triangular board and one space is open at the start.) I've represented the board with a simple array, each index being a spot and having a value of 1 if it has a peg, and 0 if it doesn't and also the set of valid moves with a 2 dimensional array that is 36, 3 (each move set contains 3 numbers; the peg you're moving, the peg it hops over, and the destination index.) So my problem is that in my recursive function, I'm using a lot of iteration to determine things like which space is open (or has a value of 0) and which move to use based on which space is open by looping through the arrays. Secondly, I don't understand how you would then backtrack with recursion, in the event that an incorrect move was made and you wanted to take that move back in order to choose a different one. This is what I have so far; the "a" array represents the board, the "b" array represents the moves, and the "c" array was my idea of a reminder as to which moves I used already. The functions used are helper functions that basically just loop through the arrays to find an empty space and corresponding move. : void solveGame(int a[15], int b[36][3], int c[15][3]){ int empSpace; int moveIndex; int count = 0; if(pegCount(a) < 2){ return; } else{ empSpace = findEmpty(a); moveIndex = chooseMove( a, b, empSpace); a[b[moveIndex][0]] = 0; a[b[moveIndex][1]] = 0; a[b[moveIndex][2]] = 1; c[count][0] = b[moveIndex][0]; c[count][1] = b[moveIndex][1]; c[count][2] = b[moveIndex][2]; solveGame(a,b,c); } }

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  • ZF Autoloader to load ancestor and requested class

    - by Pekka
    I am integrating Zend Framework into an existing application. I want to switch the application over to Zend's autoloading mechanism to replace dozens of include() statements. I have a specific requirement for the autoloading mechanism, though. Allow me to elaborate. The existing application uses a core library (independent from ZF), for example: /Core/Library/authentication.php /Core/Library/translation.php /Core/Library/messages.php this core library is to remain untouched at all times and serves a number of applications. The library contains classes like class ancestor_authentication { ... } class ancestor_translation { ... } class ancestor_messages { ... } in the application, there is also a Library directory: /App/Library/authentication.php /App/Library/translation.php /App/Library/messages.php these includes extend the ancestor classes and are the ones that actually get instantiated in the application. class authentication extends ancestor_authentication { } class translation extends ancestor_translation { } class messages extends ancestor_messages { } usually, these class definitions are empty. They simply extend their ancestors and provide the class name to instantiate. $authentication = new authentication(); The purpose of this solution is to be able to easily customize aspects of the application without having to patch the core libraries. Now, the autoloader I need would have to be aware of this structure. When an object of the class authentication is requested, the autoloader would have to: 1. load /Core/Library/authentication.php 2. load /App/Library/authentication.php My current approach would be creating a custom function, and binding that to Zend_Loader_Autoloader for a specific namespace prefix. Is there already a way to do this in Zend that I am overlooking? The accepted answer in this question kind of implies there is, but that may be just a bad choice of wording. Are there extensions to the Zend Autoloader that do this? Can you - I am new to ZF - think of an elegant way, conforming with the spirit of the framework, of extending the Autoloader with this functionality? I'm not necessary looking for a ready-made implementation, some pointers (This should be an extension to the xyz method that you would call like this...) would already be enough.

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  • Why jQuery selector can't work but getElementById works in this scenario?

    - by Stallman
    Here is the HTML: <html> <head> <script type="text/javascript" src="jquery-1.7.2.min.js"></script> <script type="text/javascript" charset="utf-8" src="jquery-1.7.2.js"></script> <script type="text/javascript" src="access.js"></script> </head> <body> <button id="trigger"></button> <img id= "testElement" style= "position: absolute; border-color: white; top:340px; left:615px;" width="34px" height= "34px" /> </body> </html> And the access.js file is: $(document).ready( function(){ $('#trigger').click(function(){ $('#testElement').src="success.png"; //THIS WON'T WORK. document.getElementById('testElement').src= "success.png"; //BUT THIS WORKS. }); }); I know that if I use $, the return object is a jQuery object. It's not the same as getElementById. But why the jQuery selector can't work here? I need the jQuery object to make more operations like "append/style"... Thanks. UPDATE Too much correct answers appear at almost the same time... Please give more explanations to let me decide who I should give the credit, thanks!!! Sorry for my poor understanding of your correct answer... I just want more detail. Are all the attribute nodes(src/width/height...) not the property of jQuery object? So does the jQuery selector only select DOM Element Node like ? Thank you! 3. List item

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  • Translate parse_git_branch function to zsh from bash (for prompt)

    - by yar
    I am using this function in Bash function parse_git_branch { git_status="$(git status 2> /dev/null)" pattern="^# On branch ([^${IFS}]*)" if [[ ! ${git_status}} =~ "working directory clean" ]]; then state="*" fi # add an else if or two here if you want to get more specific if [[ ${git_status} =~ ${pattern} ]]; then branch=${BASH_REMATCH[1]} echo "(${branch}${state})" fi } but I'm determined to use zsh. While I can use this perfectly as a shell script (even without a shebang) in my .zshrc the error is a parse error on this line if [[ ! ${git_status}}... What do I need to do to get it ready for zshell? Edit: The "actual error" I'm getting is " parse error near } and it refers to the line with the strange double }}, which works on Bash. Edit: Here's the final code, just for fun: parse_git_branch() { git_status="$(git status 2> /dev/null)" pattern="^# On branch ([^[:space:]]*)" if [[ ! ${git_status} =~ "working directory clean" ]]; then state="*" fi if [[ ${git_status} =~ ${pattern} ]]; then branch=${match[1]} echo "(${branch}${state})" fi } setopt PROMPT_SUBST PROMPT='$PR_GREEN%n@$PR_GREEN%m%u$PR_NO_COLOR:$PR_BLUE%2c$PR_NO_COLOR%(!.#.$)' RPROMPT='$PR_GREEN$(parse_git_branch)$PR_NO_COLOR' Thanks to everybody for your patience and help. Edit: The best answer has schooled us all: git status is porcelain (UI). Good scripting goes against GIT plumbing. Here's the final function: parse_git_branch() { in_wd="$(git rev-parse --is-inside-work-tree 2>/dev/null)" || return test "$in_wd" = true || return state='' git diff-index HEAD --quiet 2>/dev/null || state='*' branch="$(git symbolic-ref HEAD 2>/dev/null)" test -z "$branch" && branch='<detached-HEAD>' echo "(${branch#refs/heads/}${state})" } PROMPT='$PR_GREEN%n@$PR_GREEN%m%u$PR_NO_COLOR:$PR_BLUE%2c$PR_NO_COLOR%(!.#.$)' RPROMPT='$PR_GREEN$(parse_git_branch)$PR_NO_COLOR' Note that only the prompt is zsh-specific. In Bash it would be your prompt plus "\$(parse_git_branch)". This might be slower (more calls to GIT, but that's an empirical question) but it won't be broken by changes in GIT (they don't change the plumbing). And that is very important for a good script moving forward. Days Later: Ugh, it turns out that diff-index HEAD is NOT the same as checking status against working directory clean. So will this mean another plumbing call? I surely don't have time/expertise to write my own porcelain....

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  • How to setup Lucene/Solr for a B2B web app?

    - by Bill Paetzke
    Given: 1 database per client (business customer) 5000 clients Clients have between 2 to 2000 users (avg is ~100 users/client) 100k to 10 million records per database Users need to search those records often (it's the best way to navigate their data) Possibly relevant info: Several new clients each week (any time during business hours) Multiple web servers and database servers (users can login via any web server) Let's stay agnostic of language or sql brand, since Lucene (and Solr) have a breadth of support For Example: Joel Spolsky said in Podcast #11 that his hosted web app product, FogBugz On-Demand, uses Lucene. He has thousands of on-demand clients. And each client gets their own database. They use an index per client and store it in the client's database. I'm not sure on the details. And I'm not sure if this is a serious mod to Lucene. The Question: How would you setup Lucene search so that each client can only search within its database? How would you setup the index(es)? Where do you store the index(es)? Would you need to add a filter to all search queries? If a client cancelled, how would you delete their (part of the) index? (this may be trivial--not sure yet) Possible Solutions: Make an index for each client (database) Pro: Search is faster (than one-index-for-all method). Indices are relative to the size of the client's data. Con: I'm not sure what this entails, nor do I know if this is beyond Lucene's scope. Have a single, gigantic index with a database_name field. Always include database_name as a filter. Pro: Not sure. Maybe good for tech support or billing dept to search all databases for info. Con: Search is slower (than index-per-client method). Flawed security if query filter removed. One last thing: I would also accept an answer that uses Solr (the extension of Lucene). Perhaps it's better suited for this problem. Not sure.

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  • Webservice for uploading data: security considerations

    - by Philip Daubmeier
    Hi everyone! Im not sure about what authentification method I should use for my webservice. I've searched on SO, and found nothing that helped me. Preliminary Im building an application that uploads data from a local database to a server (running my webservice), where all records are merged and stored in a central database. I am currently binary serializing a DataTable, that holds a small fragment of the local database, where all uninteresting stuff is already filtered out. The byte[] (serialized DataTable), together with the userid and a hash of the users password is then uploaded to the webservice via SOAP. The application together with the webservice already work exactly like intended. The Problem The issue I am thinking about is now: What is if someone just sniffs the network traffic, 'steals' the users id and password hash to send his own SOAP message with modified data that corrupts my database? Options The approaches to solving that problem, I already thought of, are: Using ssl + certificates for establishing the connection: I dont really want to use ssl, I would prefer a simpler solution. After all, every information that is transfered to the webservice can be seen on the website later on. What I want to say is: there is no secret/financial/business-critical information, that has to be hidden. I think ssl would be sort of an overkill for that task. Encrypting the byte[]: I think that would be a performance killer, considering that the goal of the excercise was simply to authenticate the user. Hashing the users password together with the data: I kind of like the idea: Creating a checksum from the data, concatenating that checksum with the password-hash and hashing this whole thing again. That would assure the data was sent from this specific user, and the data wasnt modified. The actual question So, what do you think is the best approach in terms of meeting the following requirements? Rather simple solution (As it doesnt have to be super secure; no secret/business-critical information transfered) Easily implementable retrospectively (Dont want to write it all again :) ) Doesnt impact to much on performance What do you think of my prefered solution, the last one in the list above? Is there any alternative solution I didnt mention, that would fit better? You dont have to answer every question in detail. Just push me in the right direction. I very much appreciate every well-grounded opinion. Thanks in advance!

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  • current page highlights on child pages

    - by user557318
    Im trying to achieve current page highlights on wordpress similar to this site Alex Soth. I'm very nearly there with my css. At present i have current page highlights on pages, ie. home,calendar, projects. However when it come to current page highlights on child pages i have a problem. the indented child page list appears on hover when selecting a child page, but after page selection the menu reverts back to the standard pages menu with no visible child menus? unlike the link to the Alex Soth site where the extended menu stays and current page is highlighted I know that the answer will be a simple piece of css involving current_page_item and possible position:relative to obtain the menu staying visible after selection. But i can for the life of me work it out. Any ideas Ive attached my relivant pieces of css below?? thanks /* =Link Styles ------------------------------------------------------------------*/ input#submit { cursor: pointer; } input#searchsubmit { background: url(images/search.png) no-repeat center; } input#searchsubmit:hover { background: url(images/search.png) no-repeat center #3399FF !important; cursor: pointer; } .navigation a:hover, input#submit { background: #3399FF; color: #3399FF !important; } a { color: #666; } a:hover, a:hover span { color: #c11501 !important;background-color: #fae100; } .entry sup a, #main_nav .current_page_item a, #main_nav .current_page_ancestor a { color: #666 !important; } #main_nav h1.masthead a { color: #666; } #main_nav h1.masthead a:hover { border-right: none; } h2 a, #main_nav a { color: #3399FF; } img a, img a:hover { text-decoration: none; } .post a, .navigation a { font-weight: bold; color: #000; } .navigation a { background: #EEE; color: #666; font-weight: normal; padding: 3px 0px; border-radius: 0px; -webkit-border-radius: 0px; -moz-border-radius: 0px; } .post sup { font-size: 11px; color: #aaa; } .post sup a { border: 0; margin: 0; font-weight: normal; font-size: 10px; } #supplementary .post_nav ul.about_nav li a, #supplementary .post_nav ul.single_post_meta a, #supplementary ul.contact_key li a { color: #888888; border-bottom: 0; } /* =Main Menu ------------------------------------------------------------------*/ #main_nav ul.menu li { position: relative; } #main_nav ul.menu li:hover ul.sub-menu, #main_nav ul.menu li:hover ul.children { display: block; }

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  • Why does std::map operator[] create an object if the key doesn't exist?

    - by n1ck
    Hi, I'm pretty sure I already saw this question somewhere (comp.lang.c++? Google doesn't seem to find it there either) but a quick search here doesn't seem to find it so here it is: Why does the std::map operator[] create an object if the key doesn't exist? I don't know but for me this seems counter-intuitive if you compare to most other operator[] (like std::vector) where if you use it you must be sure that the index exists. I'm wondering what's the rationale for implementing this behavior in std::map. Like I said wouldn't it be more intuitive to act more like an index in a vector and crash (well undefined behavior I guess) when accessed with an invalid key? Refining my question after seeing the answers: Ok so far I got a lot of answers saying basically it's cheap so why not or things similar. I totally agree with that but why not use a dedicated function for that (I think one of the comment said that in java there is no operator[] and the function is called put)? My point is why doesn't map operator[] work like a vector? If I use operator[] on an out of range index on a vector I wouldn't like it to insert an element even if it was cheap because that probably mean an error in my code. My point is why isn't it the same thing with map. I mean, for me, using operator[] on a map would mean: i know this key already exist (for whatever reason, i just inserted it, I have redundancy somewhere, whatever). I think it would be more intuitive that way. That said what are the advantage of doing the current behavior with operator[] (and only for that, I agree that a function with the current behavior should be there, just not operator[])? Maybe it give clearer code that way? I don't know. Another answer was that it already existed that way so why not keep it but then, probably when they (the ones before stl) choose to implement it that way they found it provided an advantage or something? So my question is basically: why choose to implement it that way, meaning a somewhat lack of consistency with other operator[]. What benefit do it give? Thanks

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  • Best Practices for Working with Multiple Monitors in Visual Studio 2010

    - by Clever Human
    Now that Visual Studio 2010 has support for multiple monitors, I am curious how other people have their environments arranged. I have yet to come up with an arrangement that I am really satisfied with. The current best I have come up with for my 2 monitor system is to have all code windows detached. Then, on my primary monitor, I am able to have two code windows side by side (using the Windows 7 keyboard shortcuts WinKey+LeftArrow and WinKey+RightArrow.) On my secondary monitor I put the rest of the IDE with all of the tool windows that are normally on the bottom (errors list, find window, call stack, etc...) docked where the code windows normally go. I've also tried having all those things detached and having almost nothing in the IDE proper. The problems with this layout are: Newly opened code windows always open in the IDE, not on top of one of the detached windows. Detached code windows do not remember their exact placement from session to session (they are slightly off, having me to use the winkey + arrow key shortcut again and again for each window. There seems to be no way to have the code panes aware that they are on top of one another (IE -- multiple tabs.) The CTRL+TAB shortcut always displays on top of the IDE proper. The Code Panes are always "on top" of (children of) the IDE. So clicking on any code pane brings the IDE to the foreground, even when I care only about that code pane, and not the IDE. Other more minor issues... What would go a long way to making this better is having the code panes detach such that they are tab strips that can have other code panes docked within them. The new multi-monitor support in VS2010 is good, but it still seems really lacking. Can these issues be solved with an add-in? If so, is anyone aware of one? Is there a better way to work with the IDE on multiple monitors than what I am doing? NOTE: While this question is subjective (there is certainly no "this is the best way and that's final" answer) I'd really like to know possibly better methods of working with the IDE than what I have come up with. The intent is not to start a "mine's best" flame war.

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  • Struts2 form elements UI too rigid

    - by jscoot
    Hello, i found a similar problem like this: http://stackoverflow.com/questions/2314296/struts2-form-elements but no answer is given until now. So here i post my difficulties with a vengeance. I am using Struts2 with version 2.1.6. When I leave the input elements such as <s:textfield>, <s:textarea>, etc. of a <s:form> with the default theme, the elements are rendered as: <tr> <td class="tdLabel"><label for="firstname" class="label">Firstname:</label></td> <td><input type="text" name="firstname" id="firstname"/></td> </tr> <tr> <td class="tdLabel"><label for="lastname" class="label">Lastname:</label></td> <td><input type="text" name="lastname" id="lastname"/></td> </tr> Now if i want to add something, say an html label, between the two elements, the result gets messed up as described in the related question above. Another problem is: for the <s:checkbox> item, it is just not possible to add an extra title. For example, this tag <s:checkbox id="defaultprinter" name="defaultprinter" key="lbl.defaultprinter"/> is rendered as: <tr> <td valign="top" align="right"></td> <td valign="top" align="left"> <input type="checkbox" name="defaultprinter" value="true" checked="checked" id="defaultprinter"/> <label for="defaultprinter" class="checkboxLabel">Default Printer</label> </td> </tr> By only setting the attributes of <s:checkbox>, i can't add any text to the first <td> shown above (here it is empty!). I don't know if the above rigid UI problems can be solved or there is any workarounds somewhere. Thanks in advance.

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  • Bug in Safari: options.length = 0; not working as expected in Safari 4

    - by Stefan
    This is not a real question, but rather an answer to save some others the hassle of tracking this nasty bug down. I wasted hours finding this out. When using options.length = 0; to reset all options of a select element in safari, you can get mixed results depending on wether you have the Web Inspector open or not. If the web inspector is open you use myElement.options.length = 0; and after that query the options.length(), you might get back 1 instead of 0 (expected) but only if the Web Inspector is open (which is often the case when debugging problem like this). Workaround: Close the Web Inspector or call myElement.options.length = 0; twice like so: myElement.options.length = 0; myElement.options.length = 0; Testcase: <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01//EN" "http://www.w3.org/TR/html4/strict.dtd"> <html> <head> <title>Testcase</title> <script type="text/javascript" language="javascript" charset="utf-8"> function test(el){ var el = document.getElementById("sel"); alert("Before calling options.length=" + el.options.length); el.options.length = 0; alert("After calling options.length=" + el.options.length); } </script> </head> <body onLoad="test();"> <p> Make note of the numbers displayed in the Alert Dialog, then open Web inspector, reload this page and compare the numbers. </p> <select id="sel" multiple> <option label="a----------" value="a"></option> <option label="b----------" value="b"></option> <option label="c----------" value="c"></option> </select> </body> </html>

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  • remove/ignore float from outer div

    - by acidzombie24
    This may sound weird but i have some css which aligns mys divs. In one place i also use http://www.brunildo.org/test/img_center.html which centers images. Now i want my divs inside a larger div to go to another line if this one gets full. float: left seems to be the answer. The problem is it ruins my formatting. Including solution in the above link. I have this test code. If i remove the width and float it looks fine except it may take up too much space and not go to another line. I was thinking i could use float on an outerdiv and center the image within. However float: left is still breaking it. I am hoping there is a way to remove the float so each div does go left but the div inside centers correctly not breaking my formatting. <style type="text/css"> .wraptocenter { display: table-cell; text-align: center; vertical-align: middle; width: 200px; height: 200px; background: blue; } .wraptocenter * { vertical-align: middle; } /*\*//*/ .wraptocenter { display: block; } .wraptocenter span { display: inline-block; height: 100%; width: 1px; } /**/ div.c { background: red; overflow: hidden; min-width: 400px; max-width: 400px; } div.c div { float: left; } </style> <!--[if lt IE 8]><style> .wraptocenter span { display: inline-block; height: 100%; } </style><![endif]--> <div class="c"> <div> <div> <div class="wraptocenter"><span></span><img src="a.jpg" alt="/a.jpg"></div> <div class="wraptocenter"><span></span><img src="a.jpg" alt="/a.jpg"></div> <div class="wraptocenter"><span></span><img src="a.jpg" alt="/a.jpg"></div> </div></div></div>

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  • Overfocus in GridView

    - by chuck258
    I'm trying to implement a GridView that Focuses the next Item and "Overscrolls at the End of a List. E.g. 1 2 3 4 5 6 7 8 9 I want to scroll 1 2 3 4 5 6 ... just by pressing the right Key. Right now I can only Scroll 1 2 3 and then it stops and I have to scroll with the down Key. I already tried to set the focusViews in code (In the getView() method of my ArrayList Adapter, that fills the GridView) view.setId(position); view.setNextFocusLeftId(position-1); view.setNextFocusRightId(position+1); But that doesn't work. I found the boolean *Scroll(int direction) Methods on grepcode But theese are Package Local and I can't overwrite them. Any suggestions on how to solve this. Can I use another View and get the same Layout as a Gridview? I also set a OnFocusChangeListener to see what happens with no reaction. Edit: I just added this to my MainActivity, but now it seems to onKeyDown only get called when the GridView doesn't handle the KeyEvent (If the Last Item in a row is selected). @Override public boolean onKeyDown(int keyCode, KeyEvent event) { switch (keyCode) { case KeyEvent.KEYCODE_DPAD_LEFT: if (focusedView > 0) { mContainer.setSelection(--focusedView); Log.v("TEST", focusedView+""); } return true; case KeyEvent.KEYCODE_DPAD_RIGHT: if (focusedView < mAdapter.getCount() - 1) { mContainer.setSelection(++focusedView); Log.v("TEST", focusedView+""); } return true; } return super.onKeyDown(keyCode, event); } Edit 2: This is so f***ing stupid but works so damn fine :D @Override public boolean onKeyDown(int keyCode, KeyEvent event) { switch (keyCode) { case KeyEvent.KEYCODE_DPAD_LEFT: mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_UP, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_UP)); mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_RIGHT, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_RIGHT)); mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_RIGHT, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_RIGHT)); return true; case KeyEvent.KEYCODE_DPAD_RIGHT: mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_DOWN, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_DOWN)); mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_LEFT, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_LEFT)); mContainer.onKeyDown(KeyEvent.KEYCODE_DPAD_LEFT, new KeyEvent(KeyEvent.ACTION_DOWN, KeyEvent.KEYCODE_DPAD_LEFT)); return true; } return super.onKeyDown(keyCode, event); } I really don't want to post this as Answer, and I really don't want to have to use this Code because it is such a stupid workaround ;TLDR: Help still needed

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  • AnimationDrawable, when does it end?

    - by Syb
    I know there have been several people with the same question. Which is: How do i know when a frame by frame animation has ended? I have not had any useful answer on fora i visited. So i thought, let's see if they know at stackoverflow. But I could not sit still in the mean time, so i made a work around of this, but it does not really work the way i would like it to. here is the code: public class Main extends Activity { AnimationDrawable sybAnimation; /** Called when the activity is first created. */ @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.main); ImageView imageView = (ImageView)findViewById(R.id.ImageView01); imageView.setBackgroundResource(R.anim.testanimation); sybAnimation = (AnimationDrawable) imageView.getBackground(); imageView.post(new Starter()); } class Starter implements Runnable { public void run() { sybAnimation.start(); long totalDuration = 0; for(int i = 0; i< sybAnimation.getNumberOfFrames();i++){ totalDuration += sybAnimation.getDuration(i); } Timer timer = new Timer(); timer.schedule(new AnimationFollowUpTimerTask(R.id.ImageView01, R.anim.testanimation_reverse),totalDuration); } } class AnimationFollowUpTimerTask extends TimerTask { private int id; private int animationToRunId; public AnimationFollowUpTimerTask(int idOfImageView, int animationXML){ id = idOfImageView; animationToRunId = animationXML; } @Override public void run() { ImageView imageView = (ImageView)findViewById(id); imageView.setBackgroundResource(animationToRunId); AnimationDrawable anim = (AnimationDrawable) imageView.getBackground(); anim.start(); } } basically I make a timertask which is scheduled with the same time as the animation to take. In that run() I want to load a new animation into the imageView and start that animation, this however does not work. Does anyone know how to get this to work, or even better, have a better way to find out when an AnimationDrawable has ended its animation?

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  • MySQL server has gone away

    - by user491992
    Hello Friends, I executed this query on my MySql Server and it is giving me "MySQL server has gone away" Error.In following query my both table have more then 1000000 rows. SELECT a_tab_11_10.url as url,a_tab_11_10.c5 as 't1',a_tab_12_10.c3 as 't2' FROM a_tab_11_10 join a_tab_12_10 on (a_tab_11_10.url)=(a_tab_12_10.url) order by (a_tab_11_10.c5-a_tab_12_10.c3) desc limit 10 here is my log file but i am not getting it. Thank you @Faisal for answer and i check my log file but i am not getting it.. 110111 10:19:50 [Note] Plugin 'FEDERATED' is disabled. 110111 10:19:51 InnoDB: Started; log sequence number 0 945537221 110111 10:19:51 [Note] Event Scheduler: Loaded 0 events 110111 10:19:51 [Note] wampmysqld: ready for connections. Version: '5.1.36-community-log' socket: '' port: 3306 MySQL Community Server (GPL) 110111 12:35:42 [Note] wampmysqld: Normal shutdown 110111 12:35:43 [Note] Event Scheduler: Purging the queue. 0 events 110111 12:35:43 InnoDB: Starting shutdown... 110111 12:35:45 InnoDB: Shutdown completed; log sequence number 0 945538624 110111 12:35:45 [Warning] Forcing shutdown of 1 plugins 110111 12:35:45 [Note] wampmysqld: Shutdown complete 110111 12:36:39 [Note] Plugin 'FEDERATED' is disabled. 110111 12:36:40 InnoDB: Started; log sequence number 0 945538624 110111 12:36:40 [Note] Event Scheduler: Loaded 0 events 110111 12:36:40 [Note] wampmysqld: ready for connections. Version: '5.1.36-community-log' socket: '' port: 3306 MySQL Community Server (GPL) 110111 12:36:40 [Note] wampmysqld: Normal shutdown 110111 12:36:40 [Note] Event Scheduler: Purging the queue. 0 events 110111 12:36:40 InnoDB: Starting shutdown... 110111 12:36:42 InnoDB: Shutdown completed; log sequence number 0 945538634 110111 12:36:42 [Warning] Forcing shutdown of 1 plugins 110111 12:36:42 [Note] wampmysqld: Shutdown complete 110111 12:36:52 [Note] Plugin 'FEDERATED' is disabled. 110111 12:36:52 InnoDB: Started; log sequence number 0 945538634 110111 12:36:52 [Note] Event Scheduler: Loaded 0 events 110111 12:36:52 [Note] wampmysqld: ready for connections. Version: '5.1.36-community-log' socket: '' port: 3306 MySQL Community Server (GPL) 110111 12:37:42 [Note] wampmysqld: Normal shutdown 110111 12:37:42 [Note] Event Scheduler: Purging the queue. 0 events 110111 12:37:42 InnoDB: Starting shutdown... 110111 12:37:43 InnoDB: Shutdown completed; log sequence number 0 945538634 110111 12:37:43 [Warning] Forcing shutdown of 1 plugins 110111 12:37:43 [Note] wampmysqld: Shutdown complete 110111 12:37:46 [Note] Plugin 'FEDERATED' is disabled. 110111 12:37:46 InnoDB: Started; log sequence number 0 945538634 110111 12:37:46 [Note] Event Scheduler: Loaded 0 events 110111 12:37:46 [Note] wampmysqld: ready for connections. Version: '5.1.36-community-log' socket: '' port: 3306 MySQL Community Server (GPL)

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  • A versioning workflow for multiple similar (but not identical) deployments

    - by rs77
    I'm currently employed at a small non-tech organisation and have been given the role of coding the organisations' website. While I have enjoyed the task and have learnt much with web dev I've encountered a few issues that I'm hoping someone will be able to help with me or at least point me in the right direction on. A little background: The site I work on has subdomains that each have their own separate WordPress installation on - as this has been the easiest "backend" admin panel for the type of user who will be responsible for updating content (etc). Within the organisation I work under the Marketing Manager (MM) and I code according to his style guide and wire frames. While we have been working with only one subdomain since the beginning of the year the project has been relatively simple and straightforward. However, lately the workflow is becoming a little more complicated as our original subdomain has been copied over to the other subdomains. Each of the new subdomains receives minor edits to their stylesheets (eg. different pictures for background, slightly different colours here and there, etc). The issue: At the moment managing all the different subdomains has been "bearable", but the straw that's braking the camel's back at the moment has been the slight reversions the MM has required now that the CEO has seen the final product. The problem I'm having with reversions in stylesheets is that the CEO will one week state that he likes change "X" and then as the MM and I continue to modify the site (to now "Z"), will another week state that he wants us to change "X" to "W" but keeping most of the changes made in "Y". What I'm looking for is something that allows for: tracking file changes reverting changes made (or reverting back to 'a' from 'e' but including changes 'b' & 'c') easily upload necessary files to their respective WP-theme installation Does anything out there come close to addressing these issues? If so, what? Thanks for any help! PS - I'm learning Git at the moment and it seems to do the "tracking file changes" quite nicely. Haven't learnt about the reverting changes bit yet, though. Maybe for my final point I'm thinking of creating a shell script to automatically upload the files to their folders. Does Git do this too though? Addendum (alexbbrown) I had a similar problem: I ran a custom version of mediawiki where I installed various extensions in the versioned core (with svn). Each of the extensions required an section in the confit file, but the confit file also needed local configuration for each of several deployments. I could have implemented it using includes, but they would not be versioned; and rebasing branches each time is a chore. +50 experience points for a good answer in git.

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  • C++ Exam Question

    - by Carlucho
    I just took an exam where i was asked the following: Write the function body of each of the methods GenStrLen, InsertChar and StrReverse for the given code bellow. You must take into consideration the following; How strings are constructed in C++ The string must not overflow Insertion of character increases its length by 1 An empty string is indicated by StrLen = 0 class Strings { private: char str[80]; int StrLen; public: // Constructor Strings() { StrLen=0; }; // A function for returning the length of the string 'str' int GetStrLen(void) { }; // A function to inser a character 'ch' at the end of the string 'str' void InsertChar(char ch) { }; // A function to reverse the content of the string 'str' void StrReverse(void) { }; }; The answer I gave was something like this (see bellow). My one of problem is that used many extra variables and that makes me believe am not doing it the best possible way, and the other thing is that is not working.... class Strings { private: char str[80]; int StrLen; int index; // *** Had to add this *** public: Strings(){ StrLen=0; } int GetStrLen(void){ for (int i=0 ; str[i]!='\0' ; i++) index++; return index; // *** Here am getting a weird value, something like 1829584505306 *** } void InsertChar(char ch){ str[index] = ch; // *** Not sure if this is correct cuz I was not given int index *** } void StrRevrse(void){ GetStrLen(); char revStr[index+1]; for (int i=0 ; str[i]!='\0' ; i++){ for (int r=index ; r>0 ; r--) revStr[r] = str[i]; } } }; I would appreciate if anyone could explain me toughly what is the best way to have answered the question and why. Also how come my professor closes each class function like " }; " i thought that was only used for ending classes and constructors only. Thanks a lot for your help.

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  • GKSession sendDataToAllPeers getting invalid parameter

    - by cb4
    This is my first post and I wanted to start it by thanking the many stackoverflow contributors who will never know how much they helped me out these past several days as I worked to complete my first iOS app. I am indebted to them and to this site for giving them a place to post. To 'pay off' some of that debt, I hope this will help others as I have been helped... So, my app uses the GKPeerPickerController to make a connection to a 2nd device. Once connected, the devices can send text messages to each other. The receiving peer has the message displayed in a UIAlertView. Everything was working fine. Then I decided to experiment with locations and added code to get the current location. I convert it into latitude & longitude in degrees, minutes, and seconds and put them into one NSString. I added a button to my storyboard called 'Send Location' which, when tapped, sends the location to the connected peer. This is where I ran into the problem. Both the send text and send location methods call the sendPacket method with a NSString. sendPacket converts the string to NSData and calls sendDataToAllPeers. When I learned how to capture the error, it was "Invalid parameter for - sendDataToAllPeers:withDataMode:error:". [.....pause.....] Well, this was going to be a question but in writing all this to explain the problem, the answer just dawned on me. Did a few tests and verified it now works. The issue was not in sendDataToAllPeers, it was in the conversion of the NSString (strToSend) to NSData: packet = [strToSend dataUsingEncoding:NSASCIIStringEncoding]; Specifically, it was the degree sign character (little circle, ASCII 176). NSASCIIStringEncoding only includes ASCII characters up to 127 so don't use any above that. I'm sure there was a quicker way to find the problem, but I don't know Objective-C or Xcode's debugging facility well enough yet. Whew! Several hours to discover that little tidbit. I did learn a lot, though, and that's always a good thing!

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  • How to compare DateTime Objects while looping through a list?

    - by Taniq
    I'm trying to loop through a list (csv) containing two fields; a name and a date. There are various duplicated names and various dates in the list. I'm trying to deduce for each name in the list, where there are multiple instances of the same name, which corresponding date is the latest. I realise, from looking at another answer, that I need to use the DateTime.Compare method which is fine, but my problem is working out which date is later. Once I know this I need to produce a file with unique names and the latest date relating to it. This is my first question which makes me a newbie. EDIT: Initially I thought it would be 'ok' to set the LatestDate object to a date that wouldn't show up in my file, therefore making any later dates in the file the LatestDate. Here's my coding so far: using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.IO; namespace flybe_overwriter { class Program { static DateTime currentDate; static DateTime latestDate = new DateTime(1000,1,1); static HashSet<string> uniqueNames = new HashSet<string>(); static string indexpath = @"e:\flybe test\indexing.csv"; static string[] indexlist = File.ReadAllLines(indexpath); static StreamWriter outputfile = new StreamWriter(@"e:\flybe test\match.csv"); static void Main(string[] args) { foreach (string entry in indexlist) { uniqueNames.Add(entry.Split(',')[0]); } HashSet<string>.Enumerator fenum = new HashSet<string>.Enumerator(); fenum = uniqueNames.GetEnumerator(); while (fenum.MoveNext()) { string currentName = fenum.Current; foreach (string line in indexlist) { currentDate = new DateTime(Convert.ToInt32(line.Split(',')[1].Substring(4, 4)), Convert.ToInt32(line.Split(',')[1].Substring(2, 2)), Convert.ToInt32(line.Split(',')[1].Substring(0, 2))); if (currentName == line.Split(',')[0]) { if(DateTime.Compare(latestDate.Date, currentDate.Date) < 1) { // Console.WriteLine(currentName + " " + latestDate.ToShortDateString() + " is earlier than " + currentDate.ToShortDateString()); } else if (DateTime.Compare(latestDate.Date, currentDate.Date) > 1) { // Console.WriteLine(currentName + " " + latestDate.ToShortDateString() + " is later than " + currentDate.ToShortDateString()); } else if (DateTime.Compare(latestDate.Date, currentDate.Date) == 0) { // Console.WriteLine(currentName + " " + latestDate.ToShortDateString() + " is the same as " + currentDate.ToShortDateString()); } } } } } } } Any help appreciated. Thanks.

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  • Database design advice needed.

    - by user346271
    Hi all, I'm a lone developer for a telecoms company, and am after some database design advice from anyone with a bit of time to answer. I am inserting into one table ~2 million rows each day, these tables then get archived and compressed on a monthly basis. Each monthly table contains ~15,000,000 rows. Although this is increasing month on month. For every insert I do above I am combining the data from rows which belong together and creating another "correlated" table. This table is currently not being archived, as I need to make sure I never miss an update to the correlated table. (Hope that makes sense) Although in general this information should remain fairly static after a couple of days of processing. All of the above is working perfectly. However my company now wishes to perform some stats against this data, and these tables are getting too large to provide the results in what would be deemed a reasonable time. Even with the appropriate indexes set. So I guess after all the above my question is quite simple. Should I write a script which groups the data from my correlated table into smaller tables. Or should I store the queries result sets in something like memcache? I'm already using mysqls cache, but due to having limited control over how long the data is stored for, it's not working ideally. The main advantages I can see of using something like memcache: No blocking on my correlated table after the query has been cashed. Greater flexibility of sharing the collected data between the backend collector and front end processor. (i.e custom reports could be written in the backend and the results of these stored in the cache under a key which then gets shared with anyone who would want to see the data of this report) Redundancy and scalability if we start sharing this data with a large amount of customers. The main disadvantages I can see of using something like memcache: Data is not persistent if machine is rebooted / cache is flushed. The main advantages of using MySql Persistent data. Less code changes (although adding something like memcache is trivial anyway) The main disadvantages of using MySql Have to define table templates every time I want to store provide a new set of grouped data. Have to write a program which loops through the correlated data and fills these new tables. Potentially will still grow slower as the data continues to be filled. Apologies for quite a long question. It's helped me to write down these thoughts here anyway, and any advice/help/experience with dealing with this sort of problem would be greatly appreciated. Many thanks. Alan

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