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  • cakephp datetime insertion behaviour

    - by littlechad
    hi everyone this is a cakePHP question about datetime database insertion mismatch, i jumped in to this project while the whole thing is already built around 70%. here's what happen, every time i insert a data that contain a datetime, the inserted time doesn't match the inputted date, and the mismatch has no pattern or what ever, in some table the differences is 5 hours, while in others it could be 12 hours, 7 hours, or even 15 hours. i have traced this by investigating the controller, the model, the app_controller, everything but i don't find anything that indicate a datetime insertion rules. if the view : echo $form->input('start_date', array('label' => __l('start date')); i can't even find in the controller anything like: $this->data['current_controller']['start_date'] = $this->data['current_controller']['start_date']; when i use pr($this-data); to print the posted data, this is shown: [start_date] => Array ( [month] => 02 [day] => 16 [year] => 2011 [hour] => [min] => [meridian] => ) so i figured doing something like: $yearMonDay = $this->data['current_controller']['start_date']['year']."-"; $yearMonDay .= $this->data['current_controller']['start_date']['month']."-"; $yearMonDay .= $this->data['current_controller']['start_date']['day']; if(!empty($this->data['current_controller']['start_date']['hour'])){ $hourMinSec = $this->data['current_controller']['start_date']['hour'].":"; $hourMinSec .= $this->data['current_controller']['start_date']['min'].":"; $hourMinSec .= $this->data['current_controller']['start_date']['meridian']; }else{ $hourMinSec = "00:00:00"; } $this->data['Deal']['start_date'] = $yearMonDay." ".$hourMinSec; just to make sure the funny thing is that those posted datetime is inserted into the database with the mismatch value anyway. it's getting pretty frustrating, is there any suggestion on where else should i find the codes that define how the datetime should be inserted? or probably give me a clue on how to override those mismatched insertion rules? thanks

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  • Multiset container appears to stop sorting

    - by Sarah
    I would appreciate help debugging some strange behavior by a multiset container. Occasionally, the container appears to stop sorting. This is an infrequent error, apparent in only some simulations after a long time, and I'm short on ideas. (I'm an amateur programmer--suggestions of all kinds are welcome.) My container is a std::multiset that holds Event structs: typedef std::multiset< Event, std::less< Event > > EventPQ; with the Event structs sorted by their double time members: struct Event { public: explicit Event(double t) : time(t), eventID(), hostID(), s() {} Event(double t, int eid, int hid, int stype) : time(t), eventID( eid ), hostID( hid ), s(stype) {} bool operator < ( const Event & rhs ) const { return ( time < rhs.time ); } double time; ... }; The program iterates through periods of adding events with unordered times to EventPQ currentEvents and then pulling off events in order. Rarely, after some events have been added (with perfectly 'legal' times), events start getting executed out of order. What could make the events ever not get ordered properly? (Or what could mess up the iterator?) I have checked that all the added event times are legitimate (i.e., all exceed the current simulation time), and I have also confirmed that the error does not occur because two events happen to get scheduled for the same time. I'd love suggestions on how to work through this. The code for executing and adding events is below for the curious: double t = 0.0; double nextTimeStep = t + EPID_DELTA_T; EventPQ::iterator eventIter = currentEvents.begin(); while ( t < EPID_SIM_LENGTH ) { // Add some events to currentEvents while ( ( *eventIter ).time < nextTimeStep ) { Event thisEvent = *eventIter; t = thisEvent.time; executeEvent( thisEvent ); eventCtr++; currentEvents.erase( eventIter ); eventIter = currentEvents.begin(); } t = nextTimeStep; nextTimeStep += EPID_DELTA_T; } void Simulation::addEvent( double et, int eid, int hid, int s ) { assert( currentEvents.find( Event(et) ) == currentEvents.end() ); Event thisEvent( et, eid, hid, s ); currentEvents.insert( thisEvent ); }

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  • Visual studio 2010 colourizers, intellisense and the rest. Where to start!!

    - by Owen
    Ok, before I begin I realize that there is a lot of documentation on this subject but I have thus far failed to get even basic colourization working for VS2010. My goal is to simply get to a point where I can open a document and everything is coloured red, from here I can implement the relevant parsing logic. Here's what I have tried/found: 1) Downloaded all the relevent SDK's and such- Found the ook sample (http://code.msdn.microsoft.com/ookLanguage) - didn't build, didn't work. 2) Knowing almost nothing about MEF read through "Implementing a Language Service By Using the Managed Package Framework" - http://msdn.microsoft.com/en-us/library/bb166533(v=VS.100).aspx This was pretty much a copy and paste of all the basic stuff here, and also updating some references which were out of date with the sample see: http://social.msdn.microsoft.com/Forums/en-US/vsx/thread/a310fe67-afd2-4592-b295-3fc86fec7996 Now, I have got to a point where when running the package MEF appears to have hooked up correctly (I know this because with the debugger open I can see that the packages initialize and FDoIdle methods are being hit). When I open a file of the extension I have registered with the ProvideLanguageExtensionAttribute everything dies as if in an endless loop, yet no debug symbols hit (though they are loaded). Looking at the ook sample and the MEF examples they seem to be totally different approaches to the same problem. In the ook sample there are notions of Clasifications and Completion controllers which aren't mentioned in the MEF example. Also, they don't seem to create a Package or Language service, so I have no idea how it should work? With the MEF example, my assumption is that I need to hook into the "IScanner.ScanTokenAndProvideInfoAboutIt" to provide syntax highlighting? Which would be fine if I could ever hit this method. So my first question I guess is which approach should I be taking here? Or do they both somehow tie together? My second questions is, where can I find a basic fully working project that implements bog standard basic syntax highlighting and intellisense or VS2010? Thirdly, in the MEF example when I created a Package there were a bunch of test projects created for me. I appears that the integration tests launch the VS2010 test rig somehow, but the test fails. It would be good to write my service with tests but I have no idea what/how I can test each interaction so any references to testing Language services would be helpful. Finally, please throw any resource/book links my way that I may find useful. Cheers, Chris. N.B. Sorry I realize this is part question part rant, but I have never been so confused.

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  • Are there any modern GUI toolkits which implement a heirarchical menu buffer zone?

    - by scomar
    In Bruce Tognazzini's quiz on Fitt's Law, the question discussing the bottleneck in the hierarchical menu (as used in almost every modern desktop UI), talks about his design for the original Mac: The bottleneck is the passage between the first-level menu and the second-level menu. Users first slide the mouse pointer down to the category menu item. Then, they must carefully slide the mouse directly across (horizontally) in order to move the pointer into the secondary menu. The engineer who originally designed hierarchicals apparently had his forearm mounted on a track so that he could move it perfectly in a horizontal direction without any vertical component. Most of us, however, have our forarms mounted on a pivot we like to call our elbow. That means that moving our hand describes an arc, rather than a straight line. Demanding that pivoted people move a mouse pointer along in a straight line horizontally is just wrong. We are naturally going to slip downward even as we try to slide sideways. When we are not allowed to slip downward, the menu we're after is going to slam shut just before we get there. The Windows folks tried to overcome the pivot problem with a hack: If they see the user move down into range of the next item on the primary menu, they don't instantly close the second-level menu. Instead, they leave it open for around a half second, so, if users are really quick, they can be inaccurate but still get into the second-level menu before it slams shut. Unfortunately, people's reactions to heightened chance of error is to slow down, rather than speed up, a well-established phenomenon. Therefore, few users will ever figure out that moving faster could solve their problem. Microsoft's solution is exactly wrong. When I specified the Mac hierarchical menu algorthm in the mid-'80s, I called for a buffer zone shaped like a <, so that users could make an increasingly-greater error as they neared the hierarchical without fear of jumping to an unwanted menu. As long as the user's pointer was moving a few pixels over for every one down, on average, the menu stayed open, no matter how slow they moved. (Cancelling was still really easy; just deliberately move up or down.) This just blew me away! Such a simple idea which would result in a huge improvement in usability. I'm sure I'm not the only one who regularly has the next level of a menu slam shut because I don't move the mouse pointer in a perfectly horizontal line. So my question is: Are there any modern UI toolkits which implement this brilliant idea of a < shaped buffer zone in hierarchical menus? And if not, why not?!

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  • Cocoa's newbie question: is it possible to bind a NSTableView's selection to another tableview's selection?

    - by cocoaOverloaded
    http://img651.imageshack.us/img651/6999/modelsf.jpg Let'say, I've 2 entities in the Core Data's Model file, one being all "transactions" ever done by X company. The "transactions" entity has among other properties, a "DATE" property and a to-one relationship "COMPANY"(specifying the company with which X company has done that particular transaction). The other entity:"companies" of course contains all the companies' info ,with which X company has done transaction. The "companies" entity has a to-many relationships "TRANSACTIONS" which is an inverse relationship to "transactions" entity's "COMPANY" relationship. So within IB, I created a NSTableView(with its own NSArrayController) showing all the transactions on a particular Date (with the help of NSPredicate). Then I create another table view showing the to-many relationship "TRANSACTIONS" of the company of the selected transaction in the first table view(which shows transactions on a particular date). The 2nd table view's NSArrayController binding is like this: ** bind to: "name of the first tableview's controller", Controller Key: selection, Model Key Path:COMPANY.TRANSACTIONS(the to-many relationship in the "companies" entity)** Everythings work fine up to this moment, the 2nd tableview shows all the transactions X company has done with the company of the selected transactions in the 1st table view. But I have a group of textfields showing details of a particular transactions. Binding the these textfields with the controller of the 1st table view(the one showing transactions on a particular date) is pretty straightforward. But what I want to do are: 1/ Look up the transactions on a particular date in the first table view, select any one of them 2/ Then, check all previous transactions of the company of that transaction( selected in the first table view) from the 2nd table view 3/ Select any previous transactions and check the details of the transaction from that group of textfields So naturally I should have bind the textfields' gp to the 2nd table view's controller. But I found the default selected row in the 2nd table view(the one show all previous transactions of a company) wasn't the transaction I've selected in the 1st tableView for a particular date. Of course, i can manually select that transaction in the 2nd table view again.... So I just want to know if it's possible to have the 2nd table view automatically select the transaction according to the transaction I've selected in the 1st table view thr binding?? After hours of googling, I solved the problem by implementing the tableview Delegate protocol: - (void)tableViewSelectionDidChange:(NSNotification *)aNotification { if (["nameOf1stTableView" selectedRow] > -1) { NSArray *objsArray = ["nameOf2ndTableView'sController" arrangedObjects]; for (id obj in objsArray) { if ([[obj valueForKey:@"DATE"] isEqualToDate: ["nameOf1stTableView'sController".selection valueForKey:@"DATE"]]) { ["nameOf2ndTableView" selectRowIndexes:[NSIndexSet indexSetWithIndex:[objsArray indexOfObject:obj]] byExtendingSelection:NO]; } } } } But,this just look too cumbersome... can it be done with binding alone? Thanks in Advance,

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  • jquery load a specific #Div content from multiple html files

    - by Vikram
    Hello friends I am trying to make a Content Slider for my site. I have multiple HTML files and the structure of these files is like this: <div id="title"><h2>Title of the Slide</h2></div> <div id="image"><a href="http://mylink.com"><img src="image.jpg" width="600" height="300" alt="image"</a></div> <div id="content">Lorem Ipsum is simply dummy text of the printing and typesetting industry. Lorem Ipsum has been the industry's standard dummy text ever since the 1500s.</div> I have been trying to use the following script get content (but no success): <?php function render($position="") { ob_start(); foreach(glob("/slides/*.html") as $fileName) { $fname = basename( $fileName ); $curArr = file($fname); $slides[$fname ]['title'] = $curArr[0]; $slides[$fname ]['image'] = $curArr[1]; $slides[$fname ]['content'] = $curArr[2]; foreach($slides as $key => $value){ ?> <div id="slide-title"> <?php echo $value['title'] ?> </div> <div id="slide-content"> <?php echo $value['image'] ?> </div> <div id="slide-image"> <?php echo $value['content'] ?> </div> <?php }} ?> <?php return ob_get_clean(); } But then I came to know about a jQuery function.... (again no success) jQuery.noConflict(); (function($){ $(document).ready(function () { $('#slide-title').load('slides/slide1.html #title'); $('#slide-content').load('slides/slide1.html #content'); $('#slide-image').load('slides/slide1.html #image'); }); })(jQuery); Now My questions are..... Am I using the right syntax. How do I get the content from multiple files using jQuery. Please Note : My knowledge on Programming is almost '0'. I have just started learning it.

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  • Undefined template methods trick ?

    - by Matthieu M.
    A colleague of mine told me about a little piece of design he has used with his team that sent my mind boiling. It's a kind of traits class that they can specialize in an extremely decoupled way. I've had a hard time understanding how it could possibly work, and I am still unsure of the idea I have, so I thought I would ask for help here. We are talking g++ here, specifically the versions 3.4.2 and 4.3.2 (it seems to work with both). The idea is quite simple: 1- Define the interface // interface.h template <class T> struct Interface { void foo(); // the method is not implemented, it could not work if it was }; // // I do not think it is necessary // but they prefer free-standing methods with templates // because of the automatic argument deduction // template <class T> void foo(Interface<T>& interface) { interface.foo(); } 2- Define a class, and in the source file specialize the interface for this class (defining its methods) // special.h class Special {}; // special.cpp #include "interface.h" #include "special.h" // // Note that this specialization is not visible outside of this translation unit // template <> struct Interface<Special> { void foo() { std::cout << "Special" << std::endl; } }; 3- To use, it's simple too: // main.cpp #include "interface.h" class Special; // yes, it only costs a forward declaration // which helps much in term of dependencies int main(int argc, char* argv[]) { Interface<Special> special; foo(special); return 0; }; It's an undefined symbol if no translation unit defined a specialization of Interface for Special. Now, I would have thought this would require the export keyword, which to my knowledge has never been implemented in g++ (and only implemented once in a C++ compiler, with its authors advising anyone not to, given the time and effort it took them). I suspect it's got something to do with the linker resolving the templates methods... Do you have ever met anything like this before ? Does it conform to the standard or do you think it's a fortunate coincidence it works ? I must admit I am quite puzzled by the construct...

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  • A program where user enters a string and the program counts the instances of the letters

    - by user1865183
    This is the first C++ program I have ever written and I'm having trouble understanding the order in which operands must be put in. This is for a class, but it looks like I'm not supposed to use the homework tag. Sorry if I'm doing this wrong. This is my input // Get DNA string string st; cout << "Enter the DNA sequence to be analysed: "; cin >> st; This seems to work ok, but I thought I would include it incase this is what I'm doing wrong. This is what I have so far to check that the input is exclusively C,T,A, or G. It runs through the program and simply prints "Please enter a valid sequnce1, please enter a valid sequence2, ... ect. I'm sure I'm doing something very stupid, I just can't figure it out. // Check that the sequence is all C, T, A, G while (i <= st.size()){ if (st[i] != 'c' && st[i] != 'C' && st[i] != 'g' && st[i] != 'G' && st[i] != 't' && st[i] != 'T' && st[i] != 'a' && st[i] != 'A'); cout << "Please enter a valid sequence" << i++; else if (st[i] == c,C,G,t,T,a,A) i++; The second half of my program is to count the number of Cs and Gs in the sequence for (i < st.size() ; i++ ;); for (loop <= st.size() ; loop++;) if (st[loop] == 'c') { count_c++; } else if (st[loop] == C) { count_c++; } else if (st[loop] == g) { count_g++; } else if (st[loop] == G); { count_g++; } cout << "Number of instances of C = " << count_c; cout << "Number of instances of G = " << count_g; It seems like it's not looping, it will count 1 of one of the letters. How do I make it loop? I can't seem to put in endl; anywhere without getting an error back, although I know I'll need it somewhere. Any help or tips to point me in the right direction would be greatly appreciated - I've been working on this code for two days (this is embarrassing to admit).

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  • Linux configurations that would affect Java memory usage?

    - by wmacura
    Hi, Background: I have a set of java background workers I start as part of my webapp. I develop locally on Ubuntu 10.10 and deploy to an Ubuntu 10.04LTS server (a media temple (ve) instance). They're both running the same JVM: Sun JVM 1.6.0_22-b04. As part of the initialization script each worker is started with explicit Xmx, Xms, and XX:MaxPermGen settings. Yet somehow locally all 10 workers use 250MB, while on the server they use more than 2.7GB. I don't know how to begin to track this down. I thought the Ubuntu (and thus, kernel) version might make a difference, but I tried an old 10.04 VM and it behaves as expected. I've noticed that the machine does not seem to ever use memory for buffer or cache (according to htop), which seems a bit strange, but perhaps normal for a server? (edited) Some info: (server) root@devel:/app/axir/target# uname -a Linux devel 2.6.18-028stab069.5 #1 SMP Tue May 18 17:26:16 MSD 2010 x86_64 GNU/Linux (local) wiktor@beastie:~$ uname -a Linux beastie 2.6.35-25-generic #44-Ubuntu SMP Fri Jan 21 17:40:44 UTC 2011 x86_64 GNU/Linux (edited) Comparing PS output: (ps -eo "ppid,pid,cmd,rss,sz,vsz") PPID PID CMD RSS SZ VSZ (local) 1588 1615 java -cp axir-distribution. 25484 234382 937528 1615 1631 java -cp /home/wiktor/Code/ 83472 163059 652236 1615 1657 java -cp /home/wiktor/Code/ 70624 89135 356540 1615 1658 java -cp /home/wiktor/Code/ 37652 77625 310500 1615 1669 java -cp /home/wiktor/Code/ 38096 77733 310932 1615 1675 java -cp /home/wiktor/Code/ 37420 61395 245580 1615 1684 java -cp /home/wiktor/Code/ 38000 77736 310944 1615 1703 java -cp /home/wiktor/Code/ 39180 78060 312240 1615 1712 java -cp /home/wiktor/Code/ 38488 93882 375528 1615 1719 java -cp /home/wiktor/Code/ 38312 77874 311496 1615 1726 java -cp /home/wiktor/Code/ 38656 77958 311832 1615 1727 java -cp /home/wiktor/Code/ 78016 89429 357716 (server) 22522 23560 java -cp axir-distribution. 24860 285196 1140784 23560 23585 java -cp /app/axir/target/a 100764 161629 646516 23560 23667 java -cp /app/axir/target/a 72408 92682 370728 23560 23670 java -cp /app/axir/target/a 39948 97671 390684 23560 23674 java -cp /app/axir/target/a 40140 81586 326344 23560 23739 java -cp /app/axir/target/a 39688 81542 326168 They look very similar. In fact, the question now is why, if I add up the virtual memory usage on the server (3.2GB) does it more closely reflect 2.4GB of memory used (according to free), yet locally the virtual memory used adds up to a much more substantial 4.7GB but only actually uses ~250MB. It seems that perhaps memory isn't being shared as aggressively. (if that's even possible) Thank you for your help, Wiktor

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  • Suggest an alternative way to organize/build a database solution.

    - by Hamish Grubijan
    We are using Visual Studio 2010, but this was first conceived with VS2003. I will forward the best suggestions to my team. The current setup almost makes me vomit. It is a C# solution with most projects containing .sql files. Because we support Microsoft, Oracle, and Sybase, and so home-brewed a pre-processor, much like C preprocessor, except that substitutions are performed by a home-brewed C# program without using yacc and tools like that. #ifdefs are used for conditional macro definitions, and yeah - macros are the way this is done. A macro can expand to another macro or two, but this should eventually terminate. Only macros have #ifdef in them - the rest of the SQL-like code just uses these macros. Now, the various configurations: Debug, MNDebug, MNRelease, Release, SQL_APPLY_ALL, SQL_APPLY_MSFT, SQL_APPLY_ORACLE, SQL_APPLY_SYBASE, SQL_BUILD_OUTPUT_ALL, SQL_COMPILE, as well as 2 more. Also: Any CPU, Mixed Platforms, Win32. What drives me nuts is having to configure it correctly as well as choosing the right one out of 12 x 3 = 36 configurations as well as having to substitute database name depending on the type of database: config, main, or gateway. I am thinking that configuration should be reduced to just Debug, Release, and SQL_APPLY. Also, using 0, 1, and 2 seems so 80s ... Finally, I think my intention to build or not to build 3 types of databases for 3 types of vendors should be configured with just a tic tac toe board like: XOX OOX XXX In this case it would mean build MSFT+CONFIG, all SYBASE, and all GATEWAY. Still, the overall thing which uses a text file and a pre-processor and many configurations seems incredibly clunky. It is year 2010 now and someone out there is bound to have a very clean and/or creative tool/solution. The only pro would be that the existing collection of macros has been well tested. Have you ever had to write SQL that would work for several vendors? How did you do it? SqlVars.txt (Every one of 30 users makes a copy of a template and modifies this to suit their needs): // This is the default parameters file and should not be changed. // You can overwrite any of these parameters by copying the appropriate // section to override into SqlVars.txt and providing your own information. //Build types are 0-Config, 1-Main, 2-Gateway BUILD_TYPE=1 REMOVE_COMMENTS=1 // Login information used when applying to a Microsoft SQL server database SQL_APPLY_MSFT_version=SQL2005 SQL_APPLY_MSFT_database=msftdb SQL_APPLY_MSFT_server=ABC SQL_APPLY_MSFT_user=msftusr SQL_APPLY_MSFT_password=msftpwd // Login information used when applying to an Oracle database SQL_APPLY_ORACLE_version=ORACLE10g SQL_APPLY_ORACLE_server=oradb SQL_APPLY_ORACLE_user=orausr SQL_APPLY_ORACLE_password=orapwd // Login information used when applying to a Sybase database SQL_APPLY_SYBASE_version=SYBASE125 SQL_APPLY_SYBASE_database=sybdb SQL_APPLY_SYBASE_server=sybdb SQL_APPLY_SYBASE_user=sybusr SQL_APPLY_SYBASE_password=sybpwd ... (THIS GOES ON)

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  • jQuery Sales Tax

    - by CKallemeres
    Hello everyone! I have created a function (see below) that calculates a 7.5% sales tax. Now I need help doing the following: Have totalTax() take in 2 arguments one for the price and one for the tax. On submit (use the onSubmit event handler to call this function) have the function process the price and the tax by manipulating the arguments you passed in. Have the sales tax on the page update dynamically with what ever the sales tax is that you defined for the function 7.5 percent sales tax: Instead of using .innerHTML use jQuery to access these document elements and write to them: document.getElementById('requestedAmount' ).innerHTML = priceInput; document.getElementById('requestedTax' ).innerHTML = salesTax; document.getElementById('requestedTotal' ).innerHTML = totalAmount; Original Code: <script type="text/javascript"> $().ready(function() { // validate the comment form when it is submitted $("#inputForm").validate(); $("#priceInput").priceFormat({ prefix: '', limit: 5, centsLimit: 2 }); }); function totalTax(){ var priceInput = document.getElementById( 'priceInput' ).value; var salesTax = Math.round(((priceInput / 100) * 7.5)*100)/100; var totalAmount = (priceInput*1) + (salesTax * 1); document.getElementById( 'requestedAmount' ).innerHTML = priceInput; document.getElementById( 'requestedTax' ).innerHTML = salesTax; document.getElementById( 'requestedTotal' ).innerHTML = totalAmount; } </script> <body> <form class="cmxform" id="inputForm" method="get" action=""> <p> <label for="priceInput">Enter the price: </label> <input id="priceInput" name="name" class="required"/> </p> <p> <input class="submit" type="submit" value="Submit" onclick="totalTax();"/> </p> </form> <div>Entered price: <p id="requestedAmount"></p> </div> <div>7.5 percent sales tax: <p id="requestedTax"></p> </div> <div>Total: <p id="requestedTotal"> </p> </div>

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  • Why does OpenGL's glDrawArrays() fail with GL_INVALID_OPERATION under Core Profile 3.2, but not 3.3 or 4.2?

    - by metaleap
    I have OpenGL rendering code calling glDrawArrays that works flawlessly when the OpenGL context is (automatically / implicitly obtained) 4.2 but fails consistently (GL_INVALID_OPERATION) with an explicitly requested OpenGL core context 3.2. (Shaders are always set to #version 150 in both cases but that's beside the point here I suspect.) According to specs, there are only two instances when glDrawArrays() fails with GL_INVALID_OPERATION: "if a non-zero buffer object name is bound to an enabled array and the buffer object's data store is currently mapped" -- I'm not doing any buffer mapping at this point "if a geometry shader is active and mode? is incompatible with [...]" -- nope, no geometry shaders as of now. Furthermore: I have verified & double-checked that it's only the glDrawArrays() calls failing. Also double-checked that all arguments passed to glDrawArrays() are identical under both GL versions, buffer bindings too. This happens across 3 different nvidia GPUs and 2 different OSes (Win7 and OSX, both 64-bit -- of course, in OSX we have only the 3.2 context, no 4.2 anyway). It does not happen with an integrated "Intel HD" GPU but for that one, I only get an automatic implicit 3.3 context (trying to explicitly force a 3.2 core profile with this GPU via GLFW here fails the window creation but that's an entirely different issue...) For what it's worth, here's the relevant routine excerpted from the render loop, in Golang: func (me *TMesh) render () { curMesh = me curTechnique.OnRenderMesh() gl.BindBuffer(gl.ARRAY_BUFFER, me.glVertBuf) if me.glElemBuf > 0 { gl.BindBuffer(gl.ELEMENT_ARRAY_BUFFER, me.glElemBuf) gl.VertexAttribPointer(curProg.AttrLocs["aPos"], 3, gl.FLOAT, gl.FALSE, 0, gl.Pointer(nil)) gl.DrawElements(me.glMode, me.glNumIndices, gl.UNSIGNED_INT, gl.Pointer(nil)) gl.BindBuffer(gl.ELEMENT_ARRAY_BUFFER, 0) } else { gl.VertexAttribPointer(curProg.AttrLocs["aPos"], 3, gl.FLOAT, gl.FALSE, 0, gl.Pointer(nil)) /* BOOM! */ gl.DrawArrays(me.glMode, 0, me.glNumVerts) } gl.BindBuffer(gl.ARRAY_BUFFER, 0) } So of course this is part of a bigger render-loop, though the whole "*TMesh" construction for now is just two instances, one a simple cube and the other a simple pyramid. What matters is that the entire drawing loop works flawlessly with no errors reported when GL is queried for errors under both 3.3 and 4.2, yet on 3 nvidia GPUs with an explicit 3.2 core profile fails with an error code that according to spec is only invoked in two specific situations, none of which as far as I can tell apply here. What could be wrong here? Have you ever run into this? Any ideas what I have been missing?

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  • Is there a way for a user to disable an AlertDialog completely?

    - by NewGuyChris
    In the app I'm making, I have an "if" statement where if two strings are saved to a certain string, an AlertDialog pops up. These strings will stay the same for some users, thus having this AlertDialog constantly pop up whenever they launch the activity where the ALertDialog is set to appear. Code (I have no setNegativeButton as of yet): private void SetWarning() { AlertDialog.Builder alert = new AlertDialog.Builder(this); alert.setTitle("Warning!"); alert.setMessage(R.string.Warning); alert.setPositiveButton("Ok", new DialogInterface.OnClickListener() { public void onClick(DialogInterface dialog, int whichButton) { //No action needed; just close the AlertDialog. } }); alert.show(); } Here is a segment of my code that makes this AlertDialog appear: SharedPreferences sharedPreferences = getSharedPreferences("MY_PREF", MODE_PRIVATE); String s = sharedPreferences2.getString("MEM1", ""); String s2 = sharedPreferences2.getString("MEM2", ""); if(s.equals("String1") && s2.equals("String2")) SetWarning(); Is there a way to make an "alert.setNegativeButton" method where if the user clicks it, the AlertDialog will NEVER appear again? I'm thinking of maybe somehow implementing another SavedPreferences somehow so it saves the users selection and will then prevent the AlertDialog from ever appearing again. So far, to no luck. I've searched to find nothing, other than people asking how to disable buttons in an AlertDialog. Thank you! New updated code: alert.setNegativeButton("Cancel", new DialogInterface.OnClickListener() { public void onClick(DialogInterface dialog, int whichButton) { //set sharedpreferences boolean called DONTSHOWAGAIN to true; SharedPreferences sharedPreferences2 = getSharedPreferences("MY_PREF", MODE_PRIVATE); Boolean dontShowAgain = sharedPreferences2.getBoolean("dontShowAgain ", false); SharedPreferences.Editor ed = sharedPreferences2.edit(); ed.putBoolean("dontShowAgain", true); ed.commit(); } }); alert.show(); } private void StringWarning() { SharedPreferences sharedPreferences2 = getSharedPreferences("MY_PREF", MODE_PRIVATE); String s = sharedPreferences2.getString("MEM1", ""); String s2 = sharedPreferences2.getString("MEM2", ""); if(s.equals("String1") && s2.equals("String2")){ if(!dontShowAgain){ SetWarningExamConflict(); } }

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  • How can a button click method find out which item is selected in a ListView?

    - by Ian Bayley
    I have a single screen with a bank of buttons below a ListView. Entries on the ListView light up in orange when I scroll so I assume that are selected. When I then press the "Delete" button I want the onClickListener to remove the currently selected entry. But getSelectedItemPosition() always gives me -1. If I can't hope to use the GUI controls in this way, please give me another way of getting the same result. I have even tried setting the onClickListener of the List View to store the index before the button is pressed (in case pressing the button unselects the entry) but even that is always -1 it seems. Here's the code (without the modification which didn't work) package com.bayley; import android.app.Activity; import android.os.Bundle; import android.view.View; import android.widget.ArrayAdapter; import android.widget.Button; import android.widget.EditText; import android.widget.ListView; import java.util.ArrayList; /** * * @author p0074564 */ public class September extends Activity { /** Called when the activity is first created. */ @Override public void onCreate(Bundle icicle) { super.onCreate(icicle); setContentView(R.layout.main); final ListView myListView = (ListView) findViewById(R.id.myListView); Button addButton = (Button) findViewById(R.id.AddButton); Button deleteButton = (Button) findViewById(R.id.DeleteButton); final EditText editText = (EditText) findViewById(R.id.myEditText); final ArrayList<String> todoItems = new ArrayList<String>(); todoItems.add("Monday"); todoItems.add("Tuesday"); todoItems.add("Wednesday"); final ArrayAdapter<String> aa = new ArrayAdapter<String>(this, android.R.layout.simple_list_item_1, todoItems); myListView.setAdapter(aa); addButton.setOnClickListener(new Button.OnClickListener() { public void onClick(View v) { todoItems.add(editText.getText().toString()); aa.notifyDataSetChanged(); } }); deleteButton.setOnClickListener(new Button.OnClickListener() { public void onClick(View v) { // always returns -1 unfortunately ie nothing is ever selected int index = myListView.getSelectedItemPosition(); if (index >= 0) { todoItems.remove(index); } aa.notifyDataSetChanged(); } }); } }

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  • How to create multiple Repository object inside a Repository class using Unit Of Work?

    - by Santosh
    I am newbie to MVC3 application development, currently, we need following Application technologies as requirement MVC3 framework IOC framework – Autofac to manage object creation dynamically Moq – Unit testing Entity Framework Repository and Unit Of Work Pattern of Model class I have gone through many article to explore an basic idea about the above points but still I am little bit confused on the “Repository and Unit Of Work Pattern “. Basically what I understand Unit Of Work is a pattern which will be followed along with Repository Pattern in order to share the single DB Context among all Repository object, So here is my design : IUnitOfWork.cs public interface IUnitOfWork : IDisposable { IPermitRepository Permit_Repository{ get; } IRebateRepository Rebate_Repository { get; } IBuildingTypeRepository BuildingType_Repository { get; } IEEProjectRepository EEProject_Repository { get; } IRebateLookupRepository RebateLookup_Repository { get; } IEEProjectTypeRepository EEProjectType_Repository { get; } void Save(); } UnitOfWork.cs public class UnitOfWork : IUnitOfWork { #region Private Members private readonly CEEPMSEntities context = new CEEPMSEntities(); private IPermitRepository permit_Repository; private IRebateRepository rebate_Repository; private IBuildingTypeRepository buildingType_Repository; private IEEProjectRepository eeProject_Repository; private IRebateLookupRepository rebateLookup_Repository; private IEEProjectTypeRepository eeProjectType_Repository; #endregion #region IUnitOfWork Implemenation public IPermitRepository Permit_Repository { get { if (this.permit_Repository == null) { this.permit_Repository = new PermitRepository(context); } return permit_Repository; } } public IRebateRepository Rebate_Repository { get { if (this.rebate_Repository == null) { this.rebate_Repository = new RebateRepository(context); } return rebate_Repository; } } } PermitRepository .cs public class PermitRepository : IPermitRepository { #region Private Members private CEEPMSEntities objectContext = null; private IObjectSet<Permit> objectSet = null; #endregion #region Constructors public PermitRepository() { } public PermitRepository(CEEPMSEntities _objectContext) { this.objectContext = _objectContext; this.objectSet = objectContext.CreateObjectSet<Permit>(); } #endregion public IEnumerable<RebateViewModel> GetRebatesByPermitId(int _permitId) { // need to implment } } PermitController .cs public class PermitController : Controller { #region Private Members IUnitOfWork CEEPMSContext = null; #endregion #region Constructors public PermitController(IUnitOfWork _CEEPMSContext) { if (_CEEPMSContext == null) { throw new ArgumentNullException("Object can not be null"); } CEEPMSContext = _CEEPMSContext; } #endregion } So here I am wondering how to generate a new Repository for example “TestRepository.cs” using same pattern where I can create more then one Repository object like RebateRepository rebateRepo = new RebateRepository () AddressRepository addressRepo = new AddressRepository() because , what ever Repository object I want to create I need an object of UnitOfWork first as implmented in the PermitController class. So if I would follow the same in each individual Repository class that would again break the priciple of Unit Of Work and create multiple instance of object context. So any idea or suggestion will be highly appreciated. Thank you

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  • Learning C, would appreciate input on why this solution works.

    - by Keifer
    This is literally the first thing I've ever written in C, so please feel free to point out all it's flaws. :) My issue, however is this: if I write the program the way I feel is cleanest, I get a broken program: #include <sys/queue.h> #include <stdlib.h> #include <stdio.h> #include <string.h> /* Removed prototypes and non related code for brevity */ int main() { char *cmd = NULL; unsigned int acct = 0; int amount = 0; int done = 0; while (done==0) { scanf ("%s %u %i", cmd, &acct, &amount); if (strcmp (cmd, "exit") == 0) done = 1; else if ((strcmp (cmd, "dep") == 0) || (strcmp (cmd, "deb") == 0)) debit (acct, amount); else if ((strcmp (cmd, "wd") == 0) || (strcmp (cmd, "cred") == 0)) credit (acct, amount); else if (strcmp (cmd, "fee") == 0) service_fee(acct, amount); else printf("Invalid input!\n"); } return(0); } void credit(unsigned int acct, int amount) { } void debit(unsigned int acct, int amount) { } void service_fee(unsigned int acct, int amount) { } As it stands, the above generates no errors at compile, but gives me a segfault when ran. I can fix this by changing the program to pass cmd by reference when calling scanf and strcmp. The segfault goes away and is replaced by warnings for each use of strcmp at compile time. Despite the warnings, the affected code works. warning: passing arg 1 of 'strcmp' from incompatible pointer type As an added bonus, modifying the scanf and strcmp calls allows the program to progress far enough to execute return(0), at which point the thing crashes with an Abort trap. If I swap out return(0) for exit(0) then everything works as expected. This leaves me with two questions: why was the original program wrong? How can I fix it better than I have? The bit about needing to use exit instead of return has me especially baffled.

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  • Javascript problem when setting src for img element in FireFox - string parsing error?

    - by Kevin
    I'm having problems with image's on the page. I'm using Javascript to create the elements, and in FireFox it seems the string that I'm using to set the innerHTML is not being parsed correctly. I'll see this when the server page is requested with invalid GET variables. They look like this (from the PHP script's error handler): GET[] = Array ( [shrink] => true [file_id] => \' file_id \' [refresh] => \' now.getTime() \' ) This only happens for about 5% of requests, which is making it difficult to solve. I have been able to reproduce this myself in FireFox, and Firebug will show that the URL it is trying to fetch is: https://www.domain.com/secure/%27%20+%20image_src%20+%20%27 I read somewhere that it might be related to FireFox prefetching content (can't find it googling now), since it seems to only happen on FireFox. Disabling prefetching in about:config does prevent the problem from occurring, but I'm looking for another solution or workaround that doesn't involve end users changing their configurations. Here's the specifics and code: I have an empty table cell on an HTML page. In JQuery's $(document).ready() function for the page, I used JQuery's $.ajax() method to get some data from the server about what should be in that cell. It returns the file_id variable, which for simplicity I just set below. It then sets the empty table cell to have an image with src that points to a page that will serve the image file depending on what file_id is passed. This part of the code was JQuery originally, so I changed it to straight Javascript but that didn't help anything. //get data about image from server //this is actually done through JQuery's $.ajax() but you get the idea var file_id = 12; //create the src for the img //the refresh is to prevent the image from being cached ever, since the page's //javascript will be it changes //during the course of the page's life var now = new Date(); var image_src = 'serve_image.php?shrink=true&file_id=' + file_id + '&refresh=' + now.getTime(); //create document.getElementById('image_cell').innerHTML = '<A target="_blank" href="serve_image.php?file_id=' + file_id + '">' + '<IMG id=image_element src="' + image_src + '" alt="Loading...">' + '</A>';` Any help would be greatly appreciated. Thanks!

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  • sys.path() and PYTHONPATH issues

    - by Justin
    I've been learning Python, I'm working in 2.7.3, and I'm trying to understand import statements. The documentation says that when you attempt to import a module, the interpreter will first search for one of the built-in modules. What is meant by a built-in module? Then, the documentation says that the interpreter searches in the directories listed by sys.path, and that sys.path is initialized from these sources: the directory containing the input script (or the current directory). PYTHONPATH (a list of directory names, with the same syntax as the shell variable PATH). the installation-dependent default. Here is a sample output of a sys.path command from my computer using python in command-line mode: (I deleted a few so that it wouldn't be huge) ['', '/usr/lib/python2.7', '/usr/lib/python2.7/lib-old', '/usr/lib/python2.7/lib-dynload', '/usr/local/lib/python2.7/dist-packages', '/usr/lib/python2.7/dist-packages', '/usr/lib/python2.7/dist-packages/PIL', '/usr/lib/python2.7/dist-packages/gst-0.10', '/usr/lib/python2.7/dist-packages/gtk-2.0', '/usr/lib/pymodules/python2.7', '/usr/lib/python2.7/dist-packages/ubuntuone-couch', '/usr/lib/python2.7/dist-packages/ubuntuone-storage-protocol'] Now, I'm assuming that the '' path refers to the directory containing the 'script', and so I figured the rest of them would be coming from my PYTHONPATH environmental variable. However, when I go to the terminal and type env, PYTHONPATH doesn't exist as an environmental variable. I also tried import os then os.environ, but I get the same output. Do I really not have a PYTHONPATH environmental variable? I don't believe I ever specifically defined a PYTHONPATH environmental variable, but I assumed that when I installed new packages they automatically altered that environment variable. If I don't have a PYTHONPATH, how is my sys.path getting populated? If I download new packages, how does Python know where to look for them if I don't have this PYTHONPATH variable? How do environment variables work? From what I understand, environment variables are specific to the process for which they are set, however, if I open multiple terminal windows and run env, they all display a number of identical variables, for example, PATH. I know there file locations for persistent environment variables, for example /etc/environment, which contains my PATH variable. Is it possible to tell where a persistent environment variable is stored? What is the recommended location for storing new persistent environment variables? How do environment variables actually work with say, the Python interpreter? The Python interpreter looks for PYTHONPATH, but how does it work at the nitty-gritty level?

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  • Can't get jQuery to get focus on cloned input fields

    - by Rebel1Moon
    I have a page that needs to create dynamic form fields as often as the user needs, and I am trying to use Ajax to tie it in to my database for faster form entry and to prevent user typos. So, I have put my Ajax returned data into popup div, the user selects, then the form field is filled in. The problem comes on the cloned fields. They don't seem to want to bring up the popup div when focused. I am thinking it is something to do with when they get created/added to the DOM. Here is my JS that creates the clones: $(document).ready(function() { var regex = /^(.*)(\d)+$/i; var cloneIndex = $(".clonedInput").length; $("button.clone").live("click", function(){ $(this).parents(".clonedInput").clone() .appendTo("#course_container") .attr("id", "clonedInput" + cloneIndex) .find("*").each(function() { var id = this.id || ""; var match = id.match(regex) || []; if (match.length == 3) { this.id = match[1] + (cloneIndex); } }); cloneIndex++; numClones=cloneIndex-1; //alert("numClones "+numClones); }); Here is where I expect to be able to get focus on the correct cloned field and call the popup. The baker_equiv0 id is original code, whereas baker_equiv1 is the first clone. $('#baker_equiv0').focus(function() { \\ THIS CODE WORKS $('.popup').fadeIn(500); $('#results').empty(); // document.enter_data.baker_equiv1.value="test"; THIS LINE WORKS //alert("numClones "+numClones); }); $('#baker_equiv1').focus(function() { // THIS DOESN'T EVER FIRE alert("numClones "+numClones); $('.popup').fadeIn(500); $('#results').empty(); }); Here is the HTML with the form: <label for="baker_equiv" class="">Baker Equivalent: <span class="requiredField">*</span></label> <input type="text" class="cinputsa" name="baker_equiv[]" id="baker_equiv0" size="8" ONKEYUP="get_equiv(this.value);"> If I put this in the HTML code above, it works fine: onfocus="alert(this.id)" I'd also be interested in how to adjust the JS code to work based on the id array created rather than having to copy code for each potential set of fields clones, i.e., baker_equiv[] rather than baker_equiv0, baker_equiv1, etc. Thanks all!

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  • priority_queue with dynamic priorities

    - by Layne
    Hey, I have a server application which accepts incomming queries and executes them. If there are too many queries they should be queued and if some of the other queries got executed the queued queries should be executed as well. Since I want to pass queries with different priorities I think using a priority_queue would be the best choice. e.g. The amout of the axcepting queries (a) hit the limt and new queries will be stored in the queue. All queries have a priority of 1 (lowest) if some of the queries from (a) get executed the programm will pick the query with the highest priority out of the queue and execute it. Still no problem. Now someone is sending a query with a priority of 5 which gets added to the queue. Since this is the query with the highest priority the application will execute this query as soon as the running queries no longer hit the limit. There might be the worst case that 500 queries with a priority of 1 are queued but wont be executed since someone is always sending queries with a priority of 5 hence these 500 queries will be queued for a looooong time. In order to prevent that I want to increase the prioritiy of all queries which have a lower priority than the query with the higher priority, in this example which have a priority lower than 5. So if the query with a priority of 5 gets pulled out of the queue all other queries with a priority < 5 should be increased by 0.2. This way queries with a low priority wont be queued for ever even if there might be 100 queries with a higher priority. I really hope can help me to solve the problem with the priorities: Since my queries consist of an object I thought something like this might work: class Query { public: Query( std::string p_stQuery ) : stQuery( p_stQuery ) {}; std::string getQuery() const {return stQuery;}; void increasePriority( const float fIncrease ) {fPriority += fIncrease;}; friend bool operator < ( const Query& PriorityFirst, const Query& PriorityNext ) { if( PriorityFirst.fPriority < PriorityNext.fPriority ) { if( PriorityFirst.fStartPriority < PriorityNext.fStartPriority ) { Query qTemp = PriorityFirst; qTemp.increasePriority( INCREASE_RATE ); } return true; } else { return false; } }; private: static const float INCREASE_RATE = 0.2; float fPriority; // current priority float fStartPriority; // initialised priority std::string stQuery; };

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  • C++ Euler-Problem 14 Program Freezing

    - by Tim
    I'm working on Euler Problem 14: http://projecteuler.net/index.php?section=problems&id=14 I figured the best way would be to create a vector of numbers that kept track of how big the series was for that number... for example from 5 there are 6 steps to 1, so if ever reach the number 5 in a series, I know I have 6 steps to go and I have no need to calculate those steps. With this idea I coded up the following: #include <iostream> #include <vector> #include <iomanip> using namespace std; int main() { vector<int> sizes(1); sizes.push_back(1); sizes.push_back(2); int series, largest = 0, j; for (int i = 3; i <= 1000000; i++) { series = 0; j = i; while (j > (sizes.size()-1)) { if (j%2) { j=(3*j+1)/2; series+=2; } else { j=j/2; series++; } } series+=sizes[j]; sizes.push_back(series); if (series>largest) largest=series; cout << setw(7) << right << i << "::" << setw(5) << right << series << endl; } cout << largest << endl; return 0; } It seems to work relatively well for smaller numbers but this specific program stalls at the number 113382. Can anyone explain to me how I would go about figuring out why it freezes at this number? Is there some way I could modify my algorithim to be better? I realize that I am creating duplicates with the current way I'm doing it: for example, the series of 3 is 3,10,5,16,8,4,2,1. So I already figured out the sizes for 10,5,16,8,4,2,1 but I will duplicate those solutions later. Thanks for your help!

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  • Is it OK to put a standard, pure C header #include directive inside a namespace?

    - by mic_e
    I've got a project with a class log in the global namespace (::log). So, naturally, after #include <cmath>, the compiler gives an error message each time I try to instantiate an object of my log class, because <cmath> pollutes the global namespace with lots of three-letter methods, one of them being the logarithm function log(). So there are three possible solutions, each having their unique ugly side-effects. Move the log class to it's own namespace and always access it with it's fully qualified name. I really want to avoid this because the logger should be as convenient as possible to use. Write a mathwrapper.cpp file which is the only file in the project that includes <cmath>, and makes all the required <cmath> functions available through wrappers in a namespace math. I don't want to use this approach because I have to write a wrapper for every single required math function, and it would add additional call penalty (cancelled out partially by the -flto compiler flag) The solution I'm currently considering: Replace #include <cmath> by namespace math { #include "math.h" } and then calculating the logarithm function via math::log(). I have tried it out and it does, indeed, compile, link and run as expected. It does, however, have multiple downsides: It's (obviously) impossible to use <cmath>, because the <cmath> code accesses the functions by their fully qualified names, and it's deprecated to use in C++. I've got a really, really bad feeling about it, like I'm gonna get attacked and eaten alive by raptors. So my question is: Is there any recommendation/convention/etc that forbid putting include directives in namespaces? Could anything go wrong with diferent C standard library implementations (I use glibc), different compilers (I use g++ 4.7, -std=c++11), linking? Have you ever tried doing this? Are there any alternate ways to banish the math functions from the global namespace? I've found several similar questions on stackoverflow, but most were about including other C++ headers, which obviously is a bad idea, and those that weren't made contradictory statements about linking behaviour for C libraries. Also, would it be beneficial to additionally put the #include <math.h> inside extern "C" {}?

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  • Advice needed: stay with Java team or move to C++ team?

    - by user68759
    Some background - I have been programming in Java as a professional for the last few years. This is mainly using Java SE. I have also touched bits and pieces of other various Java technologies and have some basic knowledge about them. I consider my self as an intermediate Java programmer. I like Java very much. I think it is only going to get bigger. Recently, my manager asked my opinion on whether I would like to be transferred to another team within the company that is developing a product in C++. This is mainly because my current Java team simply didn't make enough money due to poor sales and the economic downturn. Now, I have never had any experience with C++ nor have I ever coded a single line of code in C++. I have always wanted to learn it and now is my chance. But I really want to make sure I get benefit out of it in the future, in the sense that I will have the skills that will still be on-demand in the future. So, what do you experts think? Is C++ still the language to learn these days to secure yourself for the future? What will I learn more in C++ but not in Java? And are they worthy to learn considering the current and possible future demands in IT industry? (Apart from the obvious more control over memory management and something along that line.) What is a good excuse to refuse the offer in order to stay with the Java team? I don't want to blatantly refuse it because you can never predict the future and I could possibly come back to my manager in the future and ask him to transfer me to the C++ team. How do I say it nicely that I am taking the offer but I would like to still be involved with Java one way or another, such as when there is a new Java project I would like to be considered. I have to admit that I am kind of 50-50 at the moment. I want to learn C++ for the sake of improving my skills and also helping my company to reduce the fund required for the Java team. But it is also hard for me to leave Java because I know Java is going to get bigger, so I am afraid of getting behind when I start concentrating on C++. I could, of course, decide to just join the C++ team, and then spend my free time reading about Java to keep in touch with it, but I thought I would ask anyway in case some people can point out the strong points of either over the other given the current and possibly future circumstances.

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  • What should I learn & use to become a pro in PHP & Python Web development?

    - by pecker
    Hello, I'll just show some code to show how I do web development in PHP. <html> <head> <title>Example #3 TDavid's Very First PHP Script ever!</title> </head> <? print(Date("m/j/y")); require_once("somefile.php"); $mysql_db = "DATABASE NAME"; $mysql_user = "YOUR MYSQL USERNAME"; $mysql_pass = "YOUR MYSQL PASSWORD"; $mysql_link = mysql_connect("localhost", $mysql_user, $mysql_pass); mysql_select_db($mysql_db, $mysql_link); $result = mysql_query("SELECT impressions from tds_counter where COUNT_ID='$cid'", $mysql_link); if(mysql_num_rows($result)) { mysql_query("UPDATE tds_counter set impressions=impressions+1 where COUNT_ID='$cid'", $mysql_link); $row = mysql_fetch_row($result); if(!$inv) { print("$row[0]"); } } ?> <body> </body> </html> Thats it. I write every file like this. Recently, I learnt OOP and started using classes & objects in PHP. I hear that there are many frameworks there for PHP. They say that one must use these libraries. But I feel they are just making things complicated. Anyway, this is how I've been doing my web development. Now, I want to improve this. and make it professional. Also I want to move to Python. I searched SO archives and found everyone suggesting Django. But, can any one give me some idea about how web development in Python works? user (client) request for page --- webserver(-embedded PHP interpreter) ---- Server side(PHP) Script --- MySQL Server. Now, is it that instead of PHP interpreter there is python interpreter & instead of php script there is python script, which contains both HTML & python (embedded in some kind of python tags). Python script connects to database server and fetches some data which will be printed as HTML. or is it different in python world? Is this Django thing like frameworks for PHP? Can't one code in python without using Django. Because, I never encountered any post without django Please give me some kick start.

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  • A Security (encryption) Dilemma

    - by TravisPUK
    I have an internal WPF client application that accesses a database. The application is a central resource for a Support team and as such includes Remote Access/Login information for clients. At the moment this database is not available via a web interface etc, but one day is likely to. The remote access information includes the username and passwords for the client's networks so that our client's software applications can be remotely supported by us. I need to store the usernames and passwords in the database and provide the support consultants access to them so that they can login to the client's system and then provide support. Hope this is making sense. So the dilemma is that I don't want to store the usernames and passwords in cleartext on the database to ensure that if the DB was ever compromised, I am not then providing access to our client's networks to whomever gets the database. I have looked at two-way encryption of the passwords, but as they say, two-way is not much different to cleartext as if you can decrypt it, so can an attacker... eventually. The problem here is that I have setup a method to use a salt and a passcode that are stored in the application, I have used a salt that is stored in the db, but all have their weaknesses, ie if the app was reflected it exposes the salts etc. How can I secure the usernames and passwords in my database, and yet still provide the ability for my support consultants to view the information in the application so they can use it to login? This is obviously different to storing user's passwords as these are one way because I don't need to know what they are. But I do need to know what the client's remote access passwords are as we need to enter them in at the time of remoting to them. Anybody have some theories on what would be the best approach here? update The function I am trying to build is for our CRM application that will store the remote access details for the client. The CRM system provides call/issue tracking functionality and during the course of investigating the issue, the support consultant will need to remote in. They will then view the client's remote access details and make the connection

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