Search Results

Search found 1785 results on 72 pages for 'beauness round'.

Page 67/72 | < Previous Page | 63 64 65 66 67 68 69 70 71 72  | Next Page >

  • c - dereferencing issue

    - by Joe
    Hi, I have simplified an issue that I've been having trying to isolate the problem, but it is not helping. I have a 2 dimensional char array to represent memory. I want to pass a reference to that simulation of memory to a function. In the function to test the contents of the memory I just want to iterate through the memory and print out the contents on each row. The program prints out the first row and then I get seg fault. My program is as follows: #include <stdio.h> #include <stdlib.h> #include <ctype.h> #include <string.h> void test_memory(char*** memory_ref) { int i; for(i = 0; i < 3; i++) { printf("%s\n", *memory_ref[i]); } } int main() { char** memory; int i; memory = calloc(sizeof(char*), 20); for(i = 0; i < 20; i++) { memory[i] = calloc(sizeof(char), 33); } memory[0] = "Mem 0"; memory[1] = "Mem 1"; memory[2] = "Mem 2"; printf("memory[1] = %s\n", memory[1]); test_memory(&memory); return 0; } This gives me the output: memory[1] = Mem 1 Mem 0 Segmentation fault If I change the program and create a local version of the memory in the function by dereferencing the memory_ref, then I get the right output: So: #include <stdio.h> #include <stdlib.h> #include <ctype.h> #include <string.h> void test_memory(char*** memory_ref) { char** memory = *memory_ref; int i; for(i = 0; i < 3; i++) { printf("%s\n", memory[i]); } } int main() { char** memory; int i; memory = calloc(sizeof(char*), 20); for(i = 0; i < 20; i++) { memory[i] = calloc(sizeof(char), 33); } memory[0] = "Mem 0"; memory[1] = "Mem 1"; memory[2] = "Mem 2"; printf("memory[1] = %s\n", memory[1]); test_memory(&memory); return 0; } gives me the following output: memory[1] = Mem 1 Mem 0 Mem 1 Mem 2 which is what I want, but making a local version of the memory is useless because I need to be able to change the values of the original memory from the function which I can only do by dereferencing the pointer to the original 2d char array. I don't understand why I should get a seg fault on the second time round, and I'd be grateful for any advice. Many thanks Joe

    Read the article

  • Function calls not working in my page

    - by Vivek Dragon
    I made an select menu that works with the google-font-Api. I made to function in JSBIN here is my work http://jsbin.com/ocutuk/18/ But when i made the copy of my code in a html page its not even loading the font names in page. i tried to make it work but still it is in dead end. This is my html code <!DOCTYPE html> <html> <head> <script src="http://ajax.googleapis.com/ajax/libs/jquery/1/jquery.min.js"> </script> <meta charset=utf-8 /> <title>FONT API</title> <script> function SetFonts(fonts) { for (var i = 0; i < fonts.items.length; i++) { $('#styleFont') .append($("<option></option>") .attr("value", fonts.items[i].family) .text(fonts.items[i].family)); } } var script = document.createElement('script'); script.src = 'https://www.googleapis.com/webfonts/v1/webfonts?key=AIzaSyB8Ua6XIfe-gqbkE8P3XL4spd0x8Ft7eWo&callback=SetFonts'; document.body.appendChild(script); WebFontConfig = { google: { families: ['ABeeZee', 'Abel', 'Abril Fatface', 'Aclonica', 'Acme', 'Actor', 'Adamina', 'Advent Pro', 'Aguafina Script', 'Akronim', 'Aladin', 'Aldrich', 'Alegreya', 'Alegreya SC', 'Alex Brush', 'Alfa Slab One', 'Alice', 'Alike', 'Alike Angular', 'Allan', 'Allerta', 'Allerta Stencil', 'Allura', 'Almendra', 'Almendra Display', 'Almendra SC', 'Amarante', 'Amaranth', 'Amatic SC', 'Amethysta', 'Anaheim', 'Andada', 'Andika', 'Angkor', 'Annie Use Your Telescope', 'Anonymous Pro', 'Antic', 'Antic Didone', 'Antic Slab', 'Anton', 'Arapey', 'Arbutus', 'Arbutus Slab', 'Architects Daughter', 'Archivo Black', 'Archivo Narrow', 'Arimo', 'Arizonia', 'Armata', 'Artifika', 'Arvo', 'Asap', 'Asset', 'Astloch', 'Asul', 'Atomic Age', 'Aubrey', 'Audiowide', 'Autour One', 'Average', 'Average Sans', 'Averia Gruesa Libre', 'Averia Libre', 'Averia Sans Libre', 'Averia Serif Libre', 'Bad Script', 'Balthazar', 'Bangers', 'Basic', 'Battambang', 'Baumans', 'Bayon', 'Belgrano', 'Belleza', 'BenchNine', 'Bentham', 'Berkshire Swash', 'Bevan', 'Bigelow Rules', 'Bigshot One', 'Bilbo', 'Bilbo Swash Caps', 'Bitter', 'Black Ops One', 'Bokor', 'Bonbon', 'Boogaloo', 'Bowlby One', 'Bowlby One SC', 'Brawler', 'Bree Serif', 'Bubblegum Sans', 'Bubbler One', 'Buda', 'Buenard', 'Butcherman', 'Butterfly Kids', 'Cabin', 'Cabin Condensed', 'Cabin Sketch', 'Caesar Dressing', 'Cagliostro', 'Calligraffitti', 'Cambo', 'Candal', 'Cantarell', 'Cantata One', 'Cantora One', 'Capriola', 'Cardo', 'Carme', 'Carrois Gothic', 'Carrois Gothic SC', 'Carter One', 'Caudex', 'Cedarville Cursive', 'Ceviche One', 'Changa One', 'Chango', 'Chau Philomene One', 'Chela One', 'Chelsea Market', 'Chenla', 'Cherry Cream Soda', 'Cherry Swash', 'Chewy', 'Chicle', 'Chivo', 'Cinzel', 'Cinzel Decorative', 'Clicker Script', 'Coda', 'Coda Caption', 'Codystar', 'Combo', 'Comfortaa', 'Coming Soon', 'Concert One', 'Condiment', 'Content', 'Contrail One', 'Convergence', 'Cookie', 'Copse', 'Corben', 'Courgette', 'Cousine', 'Coustard', 'Covered By Your Grace', 'Crafty Girls', 'Creepster', 'Crete Round', 'Crimson Text', 'Croissant One', 'Crushed', 'Cuprum', 'Cutive', 'Cutive Mono']} }; (function() { var wf = document.createElement('script'); wf.src = ('https:' == document.location.protocol ? 'https' : 'http') + '://ajax.googleapis.com/ajax/libs/webfont/1/webfont.js'; wf.type = 'text/javascript'; wf.async = 'true'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(wf, s); })(); $("#styleFont").change(function (){ var id =$('#styleFont option' +':selected').val(); $("#custom_text").css('font-family',id); }); </script> <style> #custom_text { font-family: Arial; resize: none; margin-top: 20px; width: 500px; } #styleFont { width: 100px; } </style> </head> <body> <select id="styleFont"> </select><br> <textarea id="custom_text"></textarea> </body> </html> How can i make it work. Whats the mistake i am making here.

    Read the article

  • PHP MINISERVER DOWNLOAD RESUME-ERROR! Resource id # 4

    - by snikolov
    $httpsock = @socket_create_listen("9090"); if (!$httpsock) { print "Socket creation failed!\n"; exit; } while (1) { $client = socket_accept($httpsock); $input = trim(socket_read ($client, 4096)); $input = explode(" ", $input); $range = $input[12]; $input = $input[1]; $fileinfo = pathinfo($input); switch ($fileinfo['extension']) { default: $mime = "text/html"; } if ($input == "/") { $input = "index.html"; } $input = ".$input"; if (file_exists($input) && is_readable($input)) { echo "Serving $input\n"; $contents = file_get_contents($input); $output = "HTTP/1.0 200 OK\r\nServer: APatchyServer\r\nConnection: close\r\nContent-Type: $mime\r\n\r\n$contents"; } else { //$contents = "The file you requested doesn't exist. Sorry!"; //$output = "HTTP/1.0 404 OBJECT NOT FOUND\r\nServer: BabyHTTP\r\nConnection: close\r\nContent-Type: text/html\r\n\r\n$contents"; if(isset($range)) { list($a, $range) = explode("=",$range); str_replace($range, "-", $range); $size2 = $size-1; $new_length = $size-$range; $output = "HTTP/1.1 206 Partial Content\r\n"; $output .= "Content-Length: $new_length\r\n"; $output .= "Content-Range: bytes $range$size2/$size\r\n"; } else { $size2=$size-1; $output .= "Content-Length: $new_length\r\n"; } $chunksize = 1*(1024*1024); $bytes_send = 0; $file = "a.mp3"; $filesize = filesize($file); if ($file = fopen($file, 'r')) { if(isset($range)) $output = 'HTTP/1.0 200 OK\r\n'; $output .= "Content-type: application/octet-stream\r\n"; $output .= "Content-Length: $filesize\r\n"; $output .= 'Content-Disposition: attachment; filename="'.$file.'"\r\n'; $output .= "Accept-Ranges: bytes\r\n"; $output .= "Cache-Control: private\n\n"; fseek($file, $range); $download_rate = 1000; while(!feof($file) and (connection_status()==0)) { $var_stat = fread($file, round($download_rate *1024)); $output .= $var_stat;//echo($buffer); // is also possible flush(); sleep(1);//// decrease download speed } fclose($file); } /** $filename = "dada"; $file = fopen($filename, 'r'); $filesize = filesize($filename); $buffer = fread($file, $filesize); $send = array("Output"=$buffer,"filesize"=$filesize,"filename"=$filename); $file = $send['filename']; */ //@ob_end_clean(); // $output .= "Content-Transfer-Encoding: binary"; //$output .= "Connection: Keep-Alive\r\n"; } socket_write($client, $output); socket_close ($client); } socket_close ($httpsock); hey guys i have create a miniwebserver downloader it can download files from your server, however i am unable to resume my download when i download the file i get Resource id # 4 and also i cant resume the download,i would like to know how i can monitor record the client output how much bandwidth he has downloaded perl has something like this put its hardcore if possible kindly provide me with some pointers thank you :)

    Read the article

  • Boost::Spirit::Qi autorules -- avoiding repeated copying of AST data structures

    - by phooji
    I've been using Qi and Karma to do some processing on several small languages. Most of the grammars are pretty small (20-40 rules). I've been able to use autorules almost exclusively, so my parse trees consist entirely of variants, structs, and std::vectors. This setup works great for the common case: 1) parse something (Qi), 2) make minor manipulations to the parse tree (visitor), and 3) output something (Karma). However, I'm concerned about what will happen if I want to make complex structural changes to a syntax tree, like moving big subtrees around. Consider the following toy example: A grammar for s-expr-style logical expressions that uses autorules... // Inside grammar class; rule names match struct names... pexpr %= pand | por | var | bconst; pand %= lit("(and ") >> (pexpr % lit(" ")) >> ")"; por %= lit("(or ") >> (pexpr % lit(" ")) >> ")"; pnot %= lit("(not ") >> pexpr >> ")"; ... which leads to parse tree representation that looks like this... struct var { std::string name; }; struct bconst { bool val; }; struct pand; struct por; struct pnot; typedef boost::variant<bconst, var, boost::recursive_wrapper<pand>, boost::recursive_wrapper<por>, boost::recursive_wrapper<pnot> > pexpr; struct pand { std::vector<pexpr> operands; }; struct por { std::vector<pexpr> operands; }; struct pnot { pexpr victim; }; // Many Fusion Macros here Suppose I have a parse tree that looks something like this: pand / ... \ por por / \ / \ var var var var (The ellipsis means 'many more children of similar shape for pand.') Now, suppose that I want negate each of the por nodes, so that the end result is: pand / ... \ pnot pnot | | por por / \ / \ var var var var The direct approach would be, for each por subtree: - create pnot node (copies por in construction); - re-assign the appropriate vector slot in the pand node (copies pnot node and its por subtree). Alternatively, I could construct a separate vector, and then replace (swap) the pand vector wholesale, eliminating a second round of copying. All of this seems cumbersome compared to a pointer-based tree representation, which would allow for the pnot nodes to be inserted without any copying of existing nodes. My question: Is there a way to avoid copy-heavy tree manipulations with autorule-compliant data structures? Should I bite the bullet and just use non-autorules to build a pointer-based AST (e.g., http://boost-spirit.com/home/2010/03/11/s-expressions-and-variants/)?

    Read the article

  • Why can't my program display this dialog box, while another program can?

    - by nonoitall
    I'm trying to write a wrapper for Winamp input plugins and have hit a bit of a snag. I'd like my wrapper to be able to display a plugin's configuration dialog, which is (or should be) achieved by calling the plugin's Config(HWND hwndParent) function. For most plugins, this works fine and my program is able to display the plugin's configuration dialog. However, 64th Note (a plugin for playing USF files) is giving me problems. Winamp can display its configuration dialog just fine, but whenever I try to display it from my wrapper, the dialog gets destroyed before it ever shows itself. Thankfully, 64th Note is open source, so I took a look at its innards to try and get an idea of what's going wrong. I've trimmed off the irrelevant bits and am left with this: Config function in the plugin (should show configuration dialog): void Config(HWND hwndParent) { DialogBox(slave, (const char *) IDD_CONFIG_WINDOW, NULL, configDlgProc); } (Slave is the plugin DLL's HINSTANCE handle.) The proc for the dialog is as follows (I have stripped out all the functionality, since it doesn't appear to have an influence on this problem): BOOL CALLBACK configDlgProc(HWND hDlg, UINT uMsg, WPARAM wParam, LPARAM lParam) { return 0; } The template for IDD_CONFIG_WINDOW is as follows: IDD_CONFIG_WINDOW DIALOGEX 0, 0, 269, 149 STYLE DS_SETFONT | DS_MODALFRAME | WS_POPUP | WS_CAPTION | WS_SYSMENU CAPTION "64th Note configuration" FONT 8, "MS Sans Serif", 0, 0, 0x0 BEGIN DEFPUSHBUTTON "OK",IDOK,212,38,50,14 CONTROL "Play Forever",IDC_NOLENGTH,"Button",BS_AUTORADIOBUTTON,7,7,55,8 CONTROL "Always Use Default Length",IDC_SETLEN,"Button",BS_AUTORADIOBUTTON,7,17,101,8 CONTROL "Default Length",IDC_DEFLEN,"Button",BS_AUTORADIOBUTTON,7,29,63,8 EDITTEXT IDC_DEFLENVAL,71,28,38,12,ES_AUTOHSCROLL EDITTEXT IDC_DEFFADEVAL,71,42,38,12,ES_AUTOHSCROLL CONTROL "Detect Silence",IDC_DETSIL,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,7,56,63,8 EDITTEXT IDC_DETSILVAL,71,56,38,12,ES_AUTOHSCROLL CONTROL "Slider2",IDC_PRISLIDER,"msctls_trackbar32",TBS_AUTOTICKS | WS_TABSTOP,74,90,108,11 EDITTEXT IDC_TITLEFMT,7,127,255,15,ES_AUTOHSCROLL CONTROL "Default to file name on missing field",IDC_FNONMISSINGTAG, "Button",BS_AUTOCHECKBOX | WS_TABSTOP,50,114,124,8 CONTROL "Use Recompiler CPU",IDC_RECOMPILER,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,7,83,8 CONTROL "Round Frequency",IDC_ROUNDFREQ,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,16,73,8 CONTROL "Seek Backwards",IDC_BACKWARDS,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,26,70,8 CONTROL "Fast Seek",IDC_FASTSEEK,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,35,48,8 CONTROL "RSP Sections",IDC_SECTIONS,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,45,60,8 CONTROL "Soft Amplify",IDC_SOFTAMP,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,54,53,8 CONTROL "Audio HLE",IDC_AUDIOHLE,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,63,50,8 CONTROL "Auto Audio HLE",IDC_AUTOAUDIOHLE,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,72,64,8 CONTROL "Display Errors",IDC_DISPERROR,"Button",BS_AUTOCHECKBOX | WS_TABSTOP,145,81,58,8 EDITTEXT IDC_RELVOL,211,104,28,12,ES_AUTOHSCROLL PUSHBUTTON "Cancel",IDCANCEL,212,54,50,14 PUSHBUTTON "Help",IDHELPBUTTON,212,71,50,14 LTEXT "Title format:",IDC_STATIC,7,113,38,8 LTEXT "seconds",IDC_STATIC,112,29,28,8 LTEXT "Default Fade",IDC_STATIC,19,43,42,8 LTEXT "seconds",IDC_STATIC,112,43,28,8 LTEXT "seconds",IDC_STATIC,112,57,28,8 CTEXT "CPU Thread Priority",IDC_STATIC,7,91,63,8 CTEXT "Look ma, I'm data!",IDC_CPUPRI,75,104,108,8 LTEXT "Relative Volume",IDC_STATIC,199,94,52,8 LTEXT "Fade Type",IDC_STATIC,7,75,35,8 COMBOBOX IDC_FADETYPE,45,72,87,74,CBS_DROPDOWNLIST | WS_TABSTOP END Naturally, without any substance in the proc function, the dialog doesn't have any functionality, but it still displays in Winamp when the Config function is invoked. However, it does not appear when I invoke it from my wrapper program. When I monitored the messages sent to the dialog in its proc function, I saw that WM_DESTROY and WM_NCDESTROY were sent within the first few messages, though I have no clue as to why. If I change the Config function so that it displays the plugin's About dialog instead of its configuration dialog, both Winamp and my wrapper will display the About dialog, which suggests that there is something unique to the configuration dialog template that's causing the problem. The modified Config function reads like so: void Config(HWND hwndParent) { DialogBox(slave, (const char *) IDD_ABOUTBOX, NULL, configDlgProc); } The template for the About dialog is as follows: IDD_ABOUTBOX DIALOGEX 0, 0, 152, 151 STYLE DS_SETFONT | DS_MODALFRAME | WS_POPUP | WS_CAPTION | WS_SYSMENU CAPTION "About 64th Note" FONT 8, "MS Sans Serif", 0, 0, 0x1 BEGIN LTEXT "64th Note v1.2 beta 3\nBased on Project 64 1.6 by Zilmar and Jabo\nAudio HLE by Azimer\nPSF concept and tagging by Neill Corlett\nPlayer by hcs, Josh W, dr0\nhttp://hcs64.com/usf",IDC_STATIC,7,94,138,50 CONTROL 110,IDC_STATIC,"Static",SS_BITMAP,26,7,95,86,WS_EX_DLGMODALFRAME END Like I said, my wrapper displays the About dialog just fine, as does Winamp. Why can Winamp display the Config dialog, while my wrapper cannot?

    Read the article

  • Delphi7 - How can i copy a file that is being written to

    - by Simon
    I have an application that logs information to a daily text file every second on a master PC. A Slave PC on the network using the same application would like to copy this text file to its local drive. I can see there is going to be file access issues. These files should be no larger than 30-40MB each. the network will be 100MB ethernet. I can see there is potential for the copying process to take longer than 1 second meaning the logging PC will need to open the file for writing while it is being read. What is the best method for the file writing(logging) and file copying procedures? I know there is the standard Windows CopyFile() procedure, however this has given me file access problems. There is also TFileStream using the fmShareDenyNone flag, but this also very occasionally gives me an access problem too (like 1 per week). What is this the best way of accomplishing this task? My current File Logging: procedure FSWriteline(Filename,Header,s : String); var LogFile : TFileStream; line : String; begin if not FileExists(filename) then begin LogFile := TFileStream.Create(FileName, fmCreate or fmShareDenyNone); try LogFile.Seek(0,soFromEnd); line := Header + #13#10; LogFile.Write(line[1],Length(line)); line := s + #13#10; LogFile.Write(line[1],Length(line)); finally logfile.Free; end; end else begin line := s + #13#10; Logfile:=tfilestream.Create(Filename,fmOpenWrite or fmShareDenyNone); try logfile.Seek(0,soFromEnd); Logfile.Write(line[1], length(line)); finally Logfile.free; end; end; end; My file copy procedure: procedure DoCopy(infile, Outfile : String); begin ForceDirectories(ExtractFilePath(outfile)); //ensure folder exists if FileAge(inFile) = FileAge(OutFile) then Exit; //they are the same modified time try { Open existing destination } fo := TFileStream.Create(Outfile, fmOpenReadWrite or fmShareDenyNone); fo.Position := 0; except { otherwise Create destination } fo := TFileStream.Create(OutFile, fmCreate or fmShareDenyNone); end; try { open source } fi := TFileStream.Create(InFile, fmOpenRead or fmShareDenyNone); try cnt:= 0; fi.Position := cnt; max := fi.Size; {start copying } Repeat dod := BLOCKSIZE; // Block size if cnt+dod>max then dod := max-cnt; if dod>0 then did := fo.CopyFrom(fi, dod); cnt:=cnt+did; Percent := Round(Cnt/Max*100); until (dod=0) finally fi.free; end; finally fo.free; end; end;

    Read the article

  • PHP mini-server download resulme-error! Resource id # 4

    - by snikolov
    <?php $httpsock = @socket_create_listen("9090"); if (!$httpsock) { print "Socket creation failed!\n"; exit; } while (1) { $client = socket_accept($httpsock); $input = trim(socket_read ($client, 4096)); $input = explode(" ", $input); $range = $input[12]; $input = $input[1]; $fileinfo = pathinfo($input); switch ($fileinfo['extension']) { default: $mime = "text/html"; } if ($input == "/") { $input = "index.html"; } $input = ".$input"; if (file_exists($input) && is_readable($input)) { echo "Serving $input\n"; $contents = file_get_contents($input); $output = "HTTP/1.0 200 OK\r\nServer: APatchyServer\r\nConnection: close\r\nContent-Type: $mime\r\n\r\n$contents"; } else { //$contents = "The file you requested doesn't exist. Sorry!"; //$output = "HTTP/1.0 404 OBJECT NOT FOUND\r\nServer: BabyHTTP\r\nConnection: close\r\nContent-Type: text/html\r\n\r\n$contents"; if(isset($range)) { list($a, $range) = explode("=",$range); str_replace($range, "-", $range); $size2 = $size-1; $new_length = $size-$range; $output = "HTTP/1.1 206 Partial Content\r\n"; $output .= "Content-Length: $new_length\r\n"; $output .= "Content-Range: bytes $range$size2/$size\r\n"; } else { $size2=$size-1; $output .= "Content-Length: $new_length\r\n"; } $chunksize = 1*(1024*1024); $bytes_send = 0; $file = "a.mp3"; $filesize = filesize($file); if ($file = fopen($file, 'r')) { if(isset($range)) $output = 'HTTP/1.0 200 OK\r\n'; $output .= "Content-type: application/octet-stream\r\n"; $output .= "Content-Length: $filesize\r\n"; $output .= 'Content-Disposition: attachment; filename="'.$file.'"\r\n'; $output .= "Accept-Ranges: bytes\r\n"; $output .= "Cache-Control: private\n\n"; fseek($file, $range); $download_rate = 1000; while(!feof($file) and (connection_status()==0)) { $var_stat = fread($file, round($download_rate *1024)); $output .= $var_stat;//echo($buffer); // is also possible flush(); sleep(1);//// decrease download speed } fclose($file); } /** $filename = "dada"; $file = fopen($filename, 'r'); $filesize = filesize($filename); $buffer = fread($file, $filesize); $send = array("Output"=>$buffer,"filesize"=>$filesize,"filename"=>$filename); $file = $send['filename']; */ //@ob_end_clean(); // $output .= "Content-Transfer-Encoding: binary"; //$output .= "Connection: Keep-Alive\r\n"; } socket_write($client, $output); socket_close ($client); } socket_close ($httpsock); Hey guys, I haved create a miniwebserver downloader. It can download files from your server. However, I am unable to resume my download when I download the file – I get Resource id # 4 – and I also can't resume the download. I would like to know how I can monitor and record the client output and how much bandwidth he has downloaded. Perl has something like this, but it's hardcore; if possible, kindly provide me with some pointers thank you :)

    Read the article

  • Android app hanging, sometimes until Force Close / Wait dialog appears

    - by fredley
    I'm making an app that records uncompressed (wav format) audio. I'm using this class to actually record the audio. Currently, my application records fine (I can play the file), however when I click the button to stop the recording, the app hangs for 10 seconds or so, with no log output or any signs of life. Finally it comes round, dumps a load of errors into the log, updates the UI etc. I'm using AsyncTasks to try and avoid this kind of thing but it's not working. Here's my code: //Called on clicks of the record button. rar is the instance of RehearsalAudioRecorder private OnClickListener RecordListener = new OnClickListener(){ @Override public void onClick(View v) { Log.d("Record","Click"); if (recording){ new stopRecordingTask().execute(rar,null,null); startStop.setText("Record"); statusBar.setText("Recording Finished, ready to Encode"); }else{ recording = true; new startRecordingTask().execute(rar,null,null); startStop.setText("Stop"); statusBar.setText("Recording Started"); } } }; private class startRecordingTask extends AsyncTask<RehearsalAudioRecorder,Void,Void>{ @Override protected Void doInBackground(RehearsalAudioRecorder... rs) { RehearsalAudioRecorder r = rs[0]; r.setOutputFile("/sdcard/rarOut.wav"); r.prepare(); r.start(); return null; } } private class stopRecordingTask extends AsyncTask<RehearsalAudioRecorder,Void,Void>{ @Override protected Void doInBackground(RehearsalAudioRecorder... rs) { RehearsalAudioRecorder r = rs[0]; r.stop(); r.reset(); return null; } } In Logcat, I always get output like this, which has me stumped. I have no idea what's causing it (I'm logging the RehearsalAudioRecorder class, and it's being started/stopped correctly by the button clicks. This output occurs after the log output for the button click and correct stop() method call) 12-19 11:59:11.172: ERROR/AudioRecord-JNI(22662): Unable to retrieve AudioRecord object, can't record 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): Error occured in updateListener, recording is aborted 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): stop() called on illegal state: STOPPED 12-19 11:59:11.172: ERROR/AudioRecord-JNI(22662): Unable to retrieve AudioRecord object, can't record 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): Error occured in updateListener, recording is aborted 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): stop() called on illegal state: ERROR 12-19 11:59:11.172: ERROR/AudioRecord-JNI(22662): Unable to retrieve AudioRecord object, can't record 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): Error occured in updateListener, recording is aborted 12-19 11:59:11.172: ERROR/uk.ac.cam.tfmw2.steg.RehearsalAudioRecorder(22662): stop() called on illegal state: ERROR ... 10 or more times I've been fiddling with this all day and I'm not getting anywhere, any input would be greatly appreciated. Update I've replace the AsyncTasks with Threads, still doesn't work, the app completely hangs when I click record, despite the fact the Log indicates there's nothing going on in the main thread. Still completely stumped.

    Read the article

  • Load balancing using Mina example with Java DSL

    - by Flame_Phoenix
    So, recently I started learning Camel. As part of the process I decided to go through all the examples (listed HERE and available when you DOWNLOAD the package with all the examples and docs) and to see what I could learn. One of the examples, Load Balancing using Mina caught my attention because it uses a Mina in different JVM's and it simulates a load balancer with round robin. I have a few problems with this example. First it uses the Spring DSL, instead of the Java DSL which my project uses and which I find a lot easier to understand now (mainly also because I am used to it). So the first question: is there a version of this example using only the Java DSL instead of the Spring DSL for the routes and the beans? My second questions is code related. The description states, and I quote: Within this demo every ten seconds, a Report object is created from the Camel load balancer server. This object is sent by the Camel load balancer to a MINA server where the object is then serialized. One of the two MINA servers (localhost:9991 and localhost:9992) receives the object and enriches the message by setting the field reply of the Report object. The reply is sent back by the MINA server to the client, which then logs the reply on the console. So, from what I read, I understand that the MINA server 1 (per example) receives a report from the loadbalancer, changes it, and then it sends that report back to some invisible client. Upon checking the code, I see no client java class or XML and when I run, the server simply posts the results on the command line. Where is the client ?? What is this client? In the MINA 1server code presented here: <beans xmlns="http://www.springframework.org/schema/beans" xmlns:xsi="http://www.w3.org/2001/XMLSchema-instance" xmlns:camel="http://camel.apache.org/schema/spring" xsi:schemaLocation=" http://www.springframework.org/schema/beans http://www.springframework.org/schema/beans/spring-beans.xsd http://camel.apache.org/schema/spring http://camel.apache.org/schema/spring/camel-spring.xsd"> <bean id="service" class="org.apache.camel.example.service.Reporting"/> <camelContext xmlns="http://camel.apache.org/schema/spring"> <route id="mina1"> <from uri="mina:tcp://localhost:9991"/> <setHeader headerName="minaServer"> <constant>localhost:9991</constant> </setHeader> <bean ref="service" method="updateReport"/> </route> </camelContext> </beans> I don't understand how the updateReport method magically prints the object on my console. What if I wanted to send message to a third MINA server? How would I do it? (I would have to add a new route, and send it to the URI of the 3rd server correct?) I know most of these questions may sound dumb, but I would appreciate if anyone could help me. A Java DSL version of this would really help me.

    Read the article

  • How to determine edges in an images optimally?

    - by SorinA.
    I recently was put in front of the problem of cropping and resizing images. I needed to crop the 'main content' of an image for example if i had an image similar to this: the result should be an image with the msn content without the white margins(left& right). I search on the X axis for the first and last color change and on the Y axis the same thing. The problem is that traversing the image line by line takes a while..for an image that is 2000x1600px it takes up to 2 seconds to return the CropRect = x1,y1,x2,y2 data. I tried to make for each coordinate a traversal and stop on the first value found but it didn't work in all test cases..sometimes the returned data wasn't the expected one and the duration of the operations was similar.. Any idea how to cut down the traversal time and discovery of the rectangle round the 'main content'? public static CropRect EdgeDetection(Bitmap Image, float Threshold) { CropRect cropRectangle = new CropRect(); int lowestX = 0; int lowestY = 0; int largestX = 0; int largestY = 0; lowestX = Image.Width; lowestY = Image.Height; //find the lowest X bound; for (int y = 0; y < Image.Height - 1; ++y) { for (int x = 0; x < Image.Width - 1; ++x) { Color currentColor = Image.GetPixel(x, y); Color tempXcolor = Image.GetPixel(x + 1, y); Color tempYColor = Image.GetPixel(x, y + 1); if ((Math.Sqrt(((currentColor.R - tempXcolor.R) * (currentColor.R - tempXcolor.R)) + ((currentColor.G - tempXcolor.G) * (currentColor.G - tempXcolor.G)) + ((currentColor.B - tempXcolor.B) * (currentColor.B - tempXcolor.B))) > Threshold)) { if (lowestX > x) lowestX = x; if (largestX < x) largestX = x; } if ((Math.Sqrt(((currentColor.R - tempYColor.R) * (currentColor.R - tempYColor.R)) + ((currentColor.G - tempYColor.G) * (currentColor.G - tempYColor.G)) + ((currentColor.B - tempYColor.B) * (currentColor.B - tempYColor.B))) > Threshold)) { if (lowestY > y) lowestY = y; if (largestY < y) largestY = y; } } } if (lowestX < Image.Width / 4) cropRectangle.X = lowestX - 3 > 0 ? lowestX - 3 : 0; else cropRectangle.X = 0; if (lowestY < Image.Height / 4) cropRectangle.Y = lowestY - 3 > 0 ? lowestY - 3 : 0; else cropRectangle.Y = 0; cropRectangle.Width = largestX - lowestX + 8 > Image.Width ? Image.Width : largestX - lowestX + 8; cropRectangle.Height = largestY + 8 > Image.Height ? Image.Height - lowestY : largestY - lowestY + 8; return cropRectangle; } }

    Read the article

  • Iphone: controlling text with delay problem with UIWebView

    - by James B.
    Hi, I've opted to use UIWebView so I can control the layout of text I've got contained in a local html document in my bundle. I want the text to display within a UIWebView I've got contained within my view. So the text isn't the whole view, just part of it. Everything runs fine, but when the web page loads I get a blank screen for a second before the text appears. This looks really bad. can anyone give me an example of how to stop this happening? I'm assuming I have to somehow hide the web view until it has fully loaded? Could someone one tell me how to do this? At the moment I'm calling my code through the viewDidLoad like this: [myUIWebView loadRequest: [NSURLRequest requestWithURL:[NSURL fileURLWithPath:[[NSBundle mainBundle]pathForResource:@"localwebpage" ofType:@"html"] isDirectory:NO]]]; Any help is much appreciated. I've read round a few forums and not seen a good answer to this question, and it seems like it recurs a lot as an issue for beginners like myself. Thanks for taking the time to read this post! UPDATED info Thanks for your response. The suggestions below solves the problem but creates a new one for me as now when my view loads it is totally hidden until I click on my toggle switch. to understand this it's maybe most helpful if I post all my code. Before this though let me explain the setup of my view. I've got a standard view within which I've also got two web views, one on top of the other. each web view contains different text with different styling. the user flicks between views using a toggle switch, which hides/reveals the web views. I'm using the web views because I want to control the style/layout of the text. Below is my full .m code, I can't figure out where it's going wrong. My web views are called oxford & harvard I'm sure its something to do with how/when I'm hiding/revealing views. I've played around with this but can't seem to get it right. Maybe my approach is wrong. A bit of advice ironing this out would be really appreciated: @implementation ReferenceViewController @synthesize oxford; @synthesize harvard; // Implement viewDidLoad to do additional setup after loading the view, typically from a nib. - (void)viewDidLoad { [super viewDidLoad]; [oxford loadRequest: [NSURLRequest requestWithURL:[NSURL fileURLWithPath:[[NSBundle mainBundle]pathForResource:@"Oxford" ofType:@"html"] isDirectory:NO]]]; [harvard loadRequest: [NSURLRequest requestWithURL:[NSURL fileURLWithPath:[[NSBundle mainBundle]pathForResource:@"Harvard" ofType:@"html"] isDirectory:NO]]]; [oxford setHidden:YES]; [harvard setHidden:YES]; } - (void)webViewDidFinishLoad:(UIWebView *)webView { if([webView hidden]) { [oxford setHidden:NO]; [harvard setHidden:NO]; } } //Toggle controls for toggle switch in UIView to swap between webviews - (IBAction)toggleControls:(id)sender { if ([sender selectedSegmentIndex] == kSwitchesSegmentIndex) { oxford.hidden = NO; harvard.hidden = YES; } else { oxford.hidden = YES; harvard.hidden = NO; } } - (void)didReceiveMemoryWarning { // Releases the view if it doesn't have a superview. [super didReceiveMemoryWarning]; // Release any cached data, images, etc that aren't in use. } - (void)viewDidUnload { [super viewDidUnload]; // Release any retained subviews of the main view. // e.g. self.myOutlet = nil; } - (void)dealloc { [super dealloc]; [oxford release]; [harvard release]; } @end

    Read the article

  • Slow Javascript touch events on Android

    - by oneself
    I'm trying to write a simple html based drawing application (standalone simplified code attached bellow). I've tested this on the following devices: iPad 1 and 2: Works great ASUS T101 running Windows: Works great Samsung Galaxy Tab: Extremely slow and patchy -- unusable. Lenovo IdeaPad K1: Extremely slow and patchy -- unusable. Asus Transformer Prime: Noticeable lag compare with the iPad -- close to usable. The Asus tablet is running ICS, the other android tablets are running 3.1 and 3.2. I tested using the stock Android browser. I also tried the Android Chrome Beta, but that was even worse. My questions is why are the Android tablets so slow? Am I doing something wrong or is it an inherit problem with Android OS or browser, or is there anything I can do about it in my code? multi.html: <html> <body> <style media="screen"> canvas { border: 1px solid #CCC; } </style> <canvas style="" id="draw" height="450" width="922"></canvas> <script class="jsbin" src="jquery.js"></script> <script src="multi.js"></script> </body> </html> multi.js: var CanvasDrawr = function(options) { // grab canvas element var canvas = document.getElementById(options.id), ctxt = canvas.getContext("2d"); canvas.style.width = '100%' canvas.width = canvas.offsetWidth; canvas.style.width = ''; // set props from options, but the defaults are for the cool kids ctxt.lineWidth = options.size || Math.ceil(Math.random() * 35); ctxt.lineCap = options.lineCap || "round"; ctxt.pX = undefined; ctxt.pY = undefined; var lines = [,,]; var offset = $(canvas).offset(); var eventCount = 0; var self = { // Bind click events init: function() { // Set pX and pY from first click canvas.addEventListener('touchstart', self.preDraw, false); canvas.addEventListener('touchmove', self.draw, false); }, preDraw: function(event) { $.each(event.touches, function(i, touch) { var id = touch.identifier; lines[id] = { x : this.pageX - offset.left, y : this.pageY - offset.top, color : 'black' }; }); event.preventDefault(); }, draw: function(event) { var e = event, hmm = {}; eventCount += 1; $.each(event.touches, function(i, touch) { var id = touch.identifier, moveX = this.pageX - offset.left - lines[id].x, moveY = this.pageY - offset.top - lines[id].y; var ret = self.move(id, moveX, moveY); lines[id].x = ret.x; lines[id].y = ret.y; }); event.preventDefault(); }, move: function(i, changeX, changeY) { ctxt.strokeStyle = lines[i].color; ctxt.beginPath(); ctxt.moveTo(lines[i].x, lines[i].y); ctxt.lineTo(lines[i].x + changeX, lines[i].y + changeY); ctxt.stroke(); ctxt.closePath(); return { x: lines[i].x + changeX, y: lines[i].y + changeY }; }, }; return self.init(); }; $(function(){ var drawr = new CanvasDrawr({ id: "draw", size: 5 }); });

    Read the article

  • can't implement jquery jScrollPane to replace browser's scrollbars

    - by Zack
    I am trying to replace browser's scrollbars with jScrollPane (jQuery), it won't work. Here are two attempts to implement it: a basic attempt, and an attempt to imitate the full page demo for jScrollPane. I've been trying everything I could think of to figure out what didn't work, but couldn't. here is my code: <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title></title> <!-- styles needed by jScrollPane --> <link type="text/css" href="style/jquery.jscrollpane.css" rel="stylesheet" media="all" /> <style type="text/css" id="page-css"> /* Styles specific to this particular page */ html { overflow: auto; } #full-page-container { overflow: auto; } .scroll-pane { width: 100%; height: 200px; overflow: auto; } .horizontal-only { height: auto; max-height: 200px; } </style> <!-- latest jQuery direct from google's CDN --> <script type="text/javascript" src="http://ajax.googleapis.com/ajax/libs/jquery/1.7.1/jquery.min.js"></script> <!-- the mousewheel plugin --> <script type="text/javascript" src="script/jquery.mousewheel.js"></script> <!-- the jScrollPane script --> <script type="text/javascript" src="script/jquery.jscrollpane.min.js"></script> <script type="text/javascript" id="sourcecode"> $(function () { var win = $(window); // Full body scroll var isResizing = false; win.bind( 'resize', function () { if (!isResizing) { isResizing = true; var container = $('#full-page-container'); // Temporarily make the container tiny so it doesn't influence the // calculation of the size of the document container.css( { 'width': 1, 'height': 1 } ); // Now make it the size of the window... container.css( { 'width': win.width(), 'height': win.height() } ); isResizing = false; container.jScrollPane( { 'showArrows': true } ); } } ).trigger('resize'); // Workaround for known Opera issue which breaks demo (see // http://jscrollpane.kelvinluck.com/known_issues.html#opera-scrollbar ) $('body').css('overflow', 'hidden'); // IE calculates the width incorrectly first time round (it // doesn't count the space used by the native scrollbar) so // we re-trigger if necessary. if ($('#full-page-container').width() != win.width()) { win.trigger('resize'); } }); </script> </head> <body> <div id="full-page-container"> This is the most basic implementation of jScrollPane I could create, if I am not wrong this has all it should take, yet it doesn't work. a little lorem ipsum to make the scrollbars show up: [here come's lot's of lorem ipsum text in the actual page...] </div> </body> </html> The other option is the same, with a link to demo.css and demo.js.

    Read the article

  • Separation of presentation and business logic in PHP

    - by Markus Ossi
    I am programming my first real PHP website and am wondering how to make my code more readable to myself. The reference book I am using is PHP and MySQL Web Development 4th ed. The aforementioned book gives three approaches to separating logic and content: include files function or class API template system I haven't chosen any of these yet, as wrapping my brains around these concepts is taking some time. However, my code has become some hybrid of the first two as I am just copy-pasting away here and modifying as I go. On presentation side, all of my pages have these common elements: header, top navigation, sidebar navigation, content, right sidebar and footer. The function-based examples in the book suggest that I could have these display functions that handle all the presentation example. So, my page code will be like this: display_header(); display_navigation(); display_content(); display_footer(); However, I don't like this because the examples in the book have these print statements with HTML and PHP mixed up like this: echo "<tr bgcolor=\"".$color."\"><td><a href=\"".$url."\">" ... I would rather like to have HTML with some PHP in the middle, not the other way round. I am thinking of making my pages so that at the beginning of my page, I will fetch all the data from database and put it in arrays. I will also get the data for variables. If there are any errors in any of these processes, I will put them into error strings. Then, at the HTML code, I will loop through these arrays using foreach and display the content. In some cases, there will be some variables that will be shown. If there is an error variable that is set, I will display that at the proper position. (As a side note: The thing I do not understand is that in most example code, if some database query or whatnot gives an error, there is always: else echo 'Error'; This baffles me, because when the example code gives an error, it is sometimes echoed out even before the HTML has started...) For people who have used ASP.NET, I have gotten somewhat used to the code-behind files and lblError and I am trying to do something similar here. The thing I haven't figured out is how could I do this "do logic first, then presentation" thing so that I would not have to replicate for example the navigation logic and navigation presentation in all of the pages. Should I do some include files or could I use functions here but a little bit differently? Are there any good articles where these "styles" of separating presentation and logic are explained a little bit more thoroughly. The book I have only has one paragraph about this stuff. What I am thinking is that I am talking about some concepts or ways of doing PHP programming here, but I just don't know the terms for them yet. I know this isn't a straight forward question, I just need some help in organizing my thoughts.

    Read the article

  • How to reduce virtual memory by optimising my PHP code?

    - by iCeR
    My current code (see below) uses 147MB of virtual memory! My provider has allocated 100MB by default and the process is killed once run, causing an internal error. The code is utilising curl multi and must be able to loop with more than 150 iterations whilst still minimizing the virtual memory. The code below is only set at 150 iterations and still causes the internal server error. At 90 iterations the issue does not occur. How can I adjust my code to lower the resource use / virtual memory? Thanks! <?php function udate($format, $utimestamp = null) { if ($utimestamp === null) $utimestamp = microtime(true); $timestamp = floor($utimestamp); $milliseconds = round(($utimestamp - $timestamp) * 1000); return date(preg_replace('`(?<!\\\\)u`', $milliseconds, $format), $timestamp); } $url = 'https://www.testdomain.com/'; $curl_arr = array(); $master = curl_multi_init(); for($i=0; $i<150; $i++) { $curl_arr[$i] = curl_init(); curl_setopt($curl_arr[$i], CURLOPT_URL, $url); curl_setopt($curl_arr[$i], CURLOPT_RETURNTRANSFER, 1); curl_setopt($curl_arr[$i], CURLOPT_SSL_VERIFYHOST, FALSE); curl_setopt($curl_arr[$i], CURLOPT_SSL_VERIFYPEER, FALSE); curl_multi_add_handle($master, $curl_arr[$i]); } do { curl_multi_exec($master,$running); } while($running > 0); for($i=0; $i<150; $i++) { $results = curl_multi_getcontent ($curl_arr[$i]); $results = explode("<br>", $results); echo $results[0]; echo "<br>"; echo $results[1]; echo "<br>"; echo udate('H:i:s:u'); echo "<br><br>"; usleep(100000); } ?> Processor Information Total processors: 8 Processor #1 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #2 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #3 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #4 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #5 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #6 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #7 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Processor #8 Vendor GenuineIntel Name Intel(R) Xeon(R) CPU E5405 @ 2.00GHz Speed 1995.120 MHz Cache 6144 KB Memory Information Memory for crash kernel (0x0 to 0x0) notwithin permissible range Memory: 8302344k/9175040k available (2176k kernel code, 80272k reserved, 901k data, 228k init, 7466304k highmem) System Information Linux server3.server.com 2.6.18-194.17.1.el5PAE #1 SMP Wed Sep 29 13:31:51 EDT 2010 i686 i686 i386 GNU/Linux Physical Disks SCSI device sda: 1952448512 512-byte hdwr sectors (999654 MB) sda: Write Protect is off sda: Mode Sense: 03 00 00 08 SCSI device sda: drive cache: write back SCSI device sda: 1952448512 512-byte hdwr sectors (999654 MB) sda: Write Protect is off sda: Mode Sense: 03 00 00 08 SCSI device sda: drive cache: write back sd 0:1:0:0: Attached scsi disk sda sd 4:0:0:0: Attached scsi removable disk sdb sd 0:1:0:0: Attached scsi generic sg4 type 0 sd 4:0:0:0: Attached scsi generic sg7 type 0 Current Memory Usage total used free shared buffers cached Mem: 8306672 7847384 459288 0 487912 6444548 -/+ buffers/cache: 914924 7391748 Swap: 4095992 496 4095496 Total: 12402664 7847880 4554784 Current Disk Usage Filesystem Size Used Avail Use% Mounted on /dev/mapper/VolGroup00-LogVol00 898G 307G 546G 36% / /dev/sda1 99M 19M 76M 20% /boot none 4.0G 0 4.0G 0% /dev/shm /var/tmpMnt 4.0G 1.8G 2.0G 48% /tmp

    Read the article

  • How to determine edges in an image optimally?

    - by SorinA.
    I recently was put in front of the problem of cropping and resizing images. I needed to crop the 'main content' of an image for example if i had an image similar to this: the result should be an image with the msn content without the white margins(left& right). I search on the X axis for the first and last color change and on the Y axis the same thing. The problem is that traversing the image line by line takes a while..for an image that is 2000x1600px it takes up to 2 seconds to return the CropRect = x1,y1,x2,y2 data. I tried to make for each coordinate a traversal and stop on the first value found but it didn't work in all test cases..sometimes the returned data wasn't the expected one and the duration of the operations was similar.. Any idea how to cut down the traversal time and discovery of the rectangle round the 'main content'? public static CropRect EdgeDetection(Bitmap Image, float Threshold) { CropRect cropRectangle = new CropRect(); int lowestX = 0; int lowestY = 0; int largestX = 0; int largestY = 0; lowestX = Image.Width; lowestY = Image.Height; //find the lowest X bound; for (int y = 0; y < Image.Height - 1; ++y) { for (int x = 0; x < Image.Width - 1; ++x) { Color currentColor = Image.GetPixel(x, y); Color tempXcolor = Image.GetPixel(x + 1, y); Color tempYColor = Image.GetPixel(x, y + 1); if ((Math.Sqrt(((currentColor.R - tempXcolor.R) * (currentColor.R - tempXcolor.R)) + ((currentColor.G - tempXcolor.G) * (currentColor.G - tempXcolor.G)) + ((currentColor.B - tempXcolor.B) * (currentColor.B - tempXcolor.B))) > Threshold)) { if (lowestX > x) lowestX = x; if (largestX < x) largestX = x; } if ((Math.Sqrt(((currentColor.R - tempYColor.R) * (currentColor.R - tempYColor.R)) + ((currentColor.G - tempYColor.G) * (currentColor.G - tempYColor.G)) + ((currentColor.B - tempYColor.B) * (currentColor.B - tempYColor.B))) > Threshold)) { if (lowestY > y) lowestY = y; if (largestY < y) largestY = y; } } } if (lowestX < Image.Width / 4) cropRectangle.X = lowestX - 3 > 0 ? lowestX - 3 : 0; else cropRectangle.X = 0; if (lowestY < Image.Height / 4) cropRectangle.Y = lowestY - 3 > 0 ? lowestY - 3 : 0; else cropRectangle.Y = 0; cropRectangle.Width = largestX - lowestX + 8 > Image.Width ? Image.Width : largestX - lowestX + 8; cropRectangle.Height = largestY + 8 > Image.Height ? Image.Height - lowestY : largestY - lowestY + 8; return cropRectangle; } }

    Read the article

  • Difficulty creating classes and arrays of those classes C#

    - by Lucifer Fayte
    I'm trying to implement a Discrete Fourier Transformation algorithm for a project I'm doing in school. But creating a class is seeming to be difficult(which it shouldn't be). I'm using Visual Studio 2012. Basically I need a class called Complex to store the two values I get from a DFT; The real portion and the imaginary portion. This is what I have so far for that: using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.Threading.Tasks; namespace SoundEditor_V3 { public class Complex { public double real; public double im; public Complex() { real = 0; im = 0; } } } The problem is that it doesn't recognize the constructor as a constructor, now I'm just learning C#, but I looked it up online and this is how it's supposed to look apparently. It recognizes my constructor as a method. Why is that? Am I creating the class wrong? It's doing the same thing for my Fourier class as well. So each time I try to create a Fourier object and then use it's method...there is no such thing. example, I do this: Fourier fou = new Fourier(); fou.DFT(s, N, amp, 0); and it tells me fou is a 'field' but is used like a 'type' why is it saying that? Here is the code for my Fourier class as well: using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.Threading.Tasks; namespace SoundEditor_V3 { public class Fourier { //FOURIER //N = number of samples //s is the array of samples(data) //amp is the array where the complex result will be written to //start is the where in the array to start public void DFT(byte[] s, int N, ref Complex[] amp, int start) { Complex tem = new Complex(); int f; int t; for (f = 0; f < N; f++) { tem.real = 0; tem.im = 0; for (t = 0; t < N; t++) { tem.real += s[t + start] * Math.Cos(2 * Math.PI * t * f / N); tem.im -= s[t + start] * Math.Sin(2 * Math.PI * t * f / N); } amp[f].real = tem.real; amp[f].im = tem.im; } } //INVERSE FOURIER public void IDFT(Complex[] A, ref int[] s) { int N = A.Length; int t, f; double result; for (t = 0; t < N; t++) { result = 0; for (f = 0; f < N; f++) { result += A[f].real * Math.Cos(2 * Math.PI * t * f / N) - A[f].im * Math.Sin(2 * Math.PI * t * f / N); } s[t] = (int)Math.Round(result); } } } } I'm very much stuck at the moment, any and all help would be appreciated. Thank you.

    Read the article

  • Flash Mouse.hide() cmd+tab(alt+tab) bring the mouse back

    - by DickieBoy
    Having a bit of trouble with Mouse.show() and losing focus of the swf This isn't my demo: but its showing the same bug http://www.foundation-flash.com/tutorials/as3customcursors/ What i do to recreate it is: mouse over the swf, hit cmd+tab to highlight another window, result is that the mouse is not brought back and is still invisible, (to get it back go to the window bar at the top of the screen and click something). I have an area in which movement is detected and an image Things I have tried package { import flash.display.Sprite; import flash.display.MovieClip; import com.greensock.*; import flash.events.*; import flash.utils.Timer; import flash.events.TimerEvent; import flash.utils.*; import flash.ui.Mouse; public class mousey_movey extends MovieClip { public var middle_of_the_flash; //pixels per second public var speeds = [0,1,3,5,10]; public var speed; public var percentage_the_mouse_is_across_the_screen; public var mouse_over_scrollable_area:Boolean; public var image_move_interval; public function mousey_movey() { middle_of_the_flash = stage.stageWidth/2; hot_area_for_movement.addEventListener(MouseEvent.MOUSE_OVER, mouseEnter); hot_area_for_movement.addEventListener(MouseEvent.MOUSE_OUT, mouseLeave); hot_area_for_movement.addEventListener(MouseEvent.MOUSE_MOVE, mouseMove); stage.addEventListener(Event.MOUSE_LEAVE, show_mouse); stage.addEventListener(MouseEvent.MOUSE_OUT, show_mouse); stage.addEventListener(Event.DEACTIVATE,show_mouse); hot_area_for_movement.alpha=0; hot_area_for_movement.x=0; hot_area_for_movement.y=34; } public function show_mouse(e) { trace(e.type) trace('show_mouse') Mouse.show(); } public function onActivate(e) { trace('activate'); Mouse.show(); } public function onDeactivate(e) { trace('deactivate'); } public function get_speed(percantage_from_middle):int { if(percantage_from_middle > 80) { return speeds[4] } else { if(percantage_from_middle > 60) { return speeds[3] } else { if(percantage_from_middle > 40) { return speeds[2] } else { if(percantage_from_middle > 20) { return speeds[1] } else { return 0; } } } } } public function mouseLeave(e:Event):void{ Mouse.show(); clearInterval(image_move_interval); } public function mouseEnter(e:Event):void{ Mouse.hide(); image_move_interval = setInterval(moveImage,1); } public function mouseMove(e:Event):void { percentage_the_mouse_is_across_the_screen = Math.round(((middle_of_the_flash-stage.mouseX)/middle_of_the_flash)*100); speed = get_speed(Math.abs(percentage_the_mouse_is_across_the_screen)); airplane_icon.x = stage.mouseX; airplane_icon.y = stage.mouseY; } public function stageMouseMove(e:Event):void{ Mouse.show(); } public function moveImage():void { if(percentage_the_mouse_is_across_the_screen > 0) { moving_image.x+=speed; airplane_icon.scaleX = -1; } else { moving_image.x-=speed; airplane_icon.scaleX = 1; } } } } Nothing too fancy, im just scrolling an image left of right at a speed which is generated by how far you are from the middle of the stage, and making an airplane moveclip follow the mouse. The events: stage.addEventListener(Event.MOUSE_LEAVE, show_mouse); stage.addEventListener(MouseEvent.MOUSE_OUT, show_mouse); stage.addEventListener(Event.DEACTIVATE,show_mouse); All fire and work correctly when in the browser, seem a little buggy when running a test through flash, was expecting this as ive experienced it before. The deactivate call even runs when testing and cmd+tabbing but shows no mouse. Any help on the matter is appreciated Thanks, Dickie

    Read the article

  • onDraw() triggered but results don't show

    - by Don
    I have the following routine in a subclass of view: It calculates an array of points that make up a line, then erases the previous lines, then draws the new lines (impact refers to the width in pixels drawn with multiple lines). The line is your basic bell curve, squeezed or stretched by variance and x-factor. Unfortunately, nothing shows on the screen. A previous version with drawPoint() and no array worked, and I've verified the array contents are being loaded correctly, and I can see that my onDraw() is being triggered. Any ideas why it might not be drawn? Thanks in advance! protected void drawNewLine( int maxx, int maxy, Canvas canvas, int impact, double variance, double xFactor, int color) { // impact = 2 to 8; xFactor between 4 and 20; variance between 0.2 and 5 double x = 0; double y = 0; int cx = maxx / 2; int cy = maxy / 2; int mu = cx; int index = 0; points[maxx<<1][1] = points[maxx<<1][0]; for (x = 0; x < maxx; x++) { points[index][1] = points[index][0]; points[index][0] = (float) x; Log.i(DEBUG_TAG, "x: " + x); index++; double root = 1.0 / (Math.sqrt(2 * Math.PI * variance)); double exponent = -1.0 * (Math.pow(((x - mu)/maxx*xFactor), 2) / (2 * variance)); double ePow = Math.exp(exponent); y = Math.round(cy * root * ePow); points[index][1] = points[index][0]; points[index][0] = (float) (maxy - y - OFFSET); index++; } points[maxx<<1][0] = (float) impact; for (int line = 0; line < points[maxx<<1][1]; line++) { for (int pt = 0; pt < (maxx<<1); pt++) { pointsToPaint[pt] = points[pt][1]; } for (int skip = 1; skip < (maxx<<1); skip = skip + 2) pointsToPaint[skip] = pointsToPaint[skip] + line; myLinePaint.setColor(Color.BLACK); canvas.drawLines(pointsToPaint, bLinePaint); // draw over old lines w/blk } for (int line = 0; line < points[maxx<<1][0]; line++) { for (int pt = 0; pt < maxx<<1; pt++) { pointsToPaint[pt] = points[pt][0]; } for (int skip = 1; skip < maxx<<1; skip = skip + 2) pointsToPaint[skip] = pointsToPaint[skip] + line; myLinePaint.setColor(color); canvas.drawLines(pointsToPaint, myLinePaint); / new color } } update: Replaced the drawLines() with drawPoint() in loop, still no joy for (int p = 0; p<pointsToPaint.length; p = p + 2) { Log.i(DEBUG_TAG, "x " + pointsToPaint[p] + " y " + pointsToPaint[p+1]); canvas.drawPoint(pointsToPaint[p], pointsToPaint[p+1], myLinePaint); } /// canvas.drawLines(pointsToPaint, myLinePaint);

    Read the article

  • Path to background in servlet

    - by kapil chhattani
    //the below line is the element of my HTML form which renders the image sent by the servlet written further below. <img style="margin-left:91px; margin-top:-6px;" class="image" src="http://www.abcd.com/captchaServlet"> I generate a captcha code using the following code in java. public class captchaServlet extends HttpServlet { protected void processRequest(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { int width = 150; int height = 50; int charsToPrint = 6; String elegibleChars = "ABCDEFGHJKLMPQRSTUVWXYabcdefhjkmnpqrstuvwxy1234567890"; char[] chars = elegibleChars.toCharArray(); StringBuffer finalString = new StringBuffer(); for ( int i = 0; i < charsToPrint; i++ ) { double randomValue = Math.random(); int randomIndex = (int) Math.round(randomValue * (chars.length - 1)); char characterToShow = chars[randomIndex]; finalString.append(characterToShow); } System.out.println(finalString); BufferedImage bufferedImage = new BufferedImage(width, height, BufferedImage.TYPE_INT_RGB); Graphics2D g2d = bufferedImage.createGraphics(); Font font = new Font("Georgia", Font.BOLD, 18); g2d.setFont(font); RenderingHints rh = new RenderingHints( RenderingHints.KEY_ANTIALIASING, RenderingHints.VALUE_ANTIALIAS_ON); rh.put(RenderingHints.KEY_RENDERING, RenderingHints.VALUE_RENDER_QUALITY); g2d.setRenderingHints(rh); GradientPaint gp = new GradientPaint(0, 0, Color.BLUE, 0, height/2, Color.black, true); g2d.setPaint(gp); g2d.fillRect(0, 0, width, height); g2d.setColor(new Color(255, 255, 0)); Random r = new Random(); int index = Math.abs(r.nextInt()) % 5; char[] data=new String(finalString).toCharArray(); String captcha = String.copyValueOf(data); int x = 0; int y = 0; for (int i=0; i<data.length; i++) { x += 10 + (Math.abs(r.nextInt()) % 15); y = 20 + Math.abs(r.nextInt()) % 20; g2d.drawChars(data, i, 1, x, y); } g2d.dispose(); response.setContentType("image/png"); OutputStream os = response.getOutputStream(); ImageIO.write(bufferedImage, "png", os); os.close(); } protected void doGet(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { processRequest(request, response); } protected void doPost(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { processRequest(request, response); } } But in the above code background is also generated using the setPaint menthod I am guessing. I want the background to be some image from my local machine whoz URL i should be able to mention like URL url=this.getClass().getResource("Desktop/images.jpg"); BufferedImage bufferedImage = ImageIO.read(url); I am just writing the above two lines for making the reader understand better what the issue is. Dont want to use the exact same commands. All I want is the the background of the captcha code generated should be an image of my choice.

    Read the article

  • CSS optimization - extra classes in dom or preprocessor-repetitive styling in css file?

    - by anna.mi
    I'm starting on a fairly large project and I'm considering the option of using LESS for pre-processing my css. the useful thing about LESS is that you can define a mixin that contains for example: .border-radius(@radius) { -webkit-border-radius: @radius; -moz-border-radius: @radius; -o-border-radius: @radius; -ms-border-radius: @radius; border-radius: @radius; } and then use it in a class declaration as .rounded-div { .border-radius(10px); } to get the outputted css as: .rounded-div { -webkit-border-radius: 10px; -moz-border-radius: 10px; -o-border-radius: 10px; -ms-border-radius: 10px; border-radius: 10px; } this is extremely useful in the case of browser prefixes. However this same concept could be used to encapsulate commonly-used css, for example: .column-container { overflow: hidden; display: block; width: 100%; } .column(@width) { float: left; width: @width; } and then use this mixin whenever i need columns in my design: .my-column-outer { .column-container(); background: red; } .my-column-inner { .column(50%); font-color: yellow; } (of course, using the preprocessor we could easily expand this to be much more useful, eg. pass the number of columns and the container width as variables and have LESS determine the width of each column depending on the number of columns and container width!) the problem with this is that when compliled, my final css file would have 100 such declarations, copy&pasted, making the file huge and bloated and repetitive. The alternative to this would be to use a grid system which has predefined classes for each column-layout option, eg .c-50 ( with a "float: left; width:50%;" definition ), .c-33, .c-25 to accomodate for a 2-column, 3-column and 4-column layout and then use these classes to my dom. i really mislike the idea of the extra classes, from experience it results to bloated dom (creating extra divs just to attach the grid classes to). Also the most basic tutorial for html/css would tell you that the dom should be separated from the styling - grid classes are styling related! to me, its the same as attaching a "border-radius-10" class to the .rounded-div example above! on the other hand, the large css file that would result from the repetitive code is also a disadvantage so i guess my question is, which one would you recommend? and which do you use? and, which solution is best for optimization? apart from the larger file size, has there even been any research on whether browser renders multiple classes faster than a large css file, or the other way round? tnx! i'd love to hear your opinion!

    Read the article

  • Suggestions for lightweight, thread-safe scheduler

    - by nirvanai
    I am trying to write a round-robin scheduler for lightweight threads (fibers). It must scale to handle as many concurrently-scheduled fibers as possible. I also need to be able to schedule fibers from threads other than the one the run loop is on, and preferably unschedule them from arbitrary threads as well (though I could live with only being able to unschedule them from the run loop). My current idea is to have a circular doubly-linked list, where each fiber is a node and the scheduler holds a reference to the current node. This is what I have so far: using Interlocked = System.Threading.Interlocked; public class Thread { internal Future current_fiber; public void RunLoop () { while (true) { var fiber = current_fiber; if (fiber == null) { // block the thread until a fiber is scheduled continue; } if (fiber.Fulfilled) fiber.Unschedule (); else fiber.Resume (); //if (current_fiber == fiber) current_fiber = fiber.next; Interlocked.CompareExchange<Future> (ref current_fiber, fiber.next, fiber); } } } public abstract class Future { public bool Fulfilled { get; protected set; } internal Future previous, next; // this must be thread-safe // it inserts this node before thread.current_fiber // (getting the exact position doesn't matter, as long as the // chosen nodes haven't been unscheduled) public void Schedule (Thread thread) { next = this; // maintain circularity, even if this is the only node previous = this; try_again: var current = Interlocked.CompareExchange<Future> (ref thread.current_fiber, this, null); if (current == null) return; var target = current.previous; while (target == null) { // current was unscheduled; negotiate for new current_fiber var potential = current.next; var actual = Interlocked.CompareExchange<Future> (ref thread.current_fiber, potential, current); current = (actual == current? potential : actual); if (current == null) goto try_again; target = current.previous; } // I would lock "current" and "target" at this point. // How can I do this w/o risk of deadlock? next = current; previous = target; target.next = this; current.previous = this; } // this would ideally be thread-safe public void Unschedule () { var prev = previous; if (prev == null) { // already unscheduled return; } previous = null; if (next == this) { next = null; return; } // Again, I would lock "prev" and "next" here // How can I do this w/o risk of deadlock? prev.next = next; next.previous = prev; } public abstract void Resume (); } As you can see, my sticking point is that I cannot ensure the order of locking, so I can't lock more than one node without risking deadlock. Or can I? I don't want to have a global lock on the Thread object, since the amount of lock contention would be extreme. Plus, I don't especially care about insertion position, so if I lock each node separately then Schedule() could use something like Monitor.TryEnter and just keep walking the list until it finds an unlocked node. Overall, I'm not invested in any particular implementation, as long as it meets the requirements I've mentioned. Any ideas would be greatly appreciated. Thanks! P.S- For the curious, this is for an open source project I'm starting at http://github.com/nirvanai/Cirrus

    Read the article

  • Entity Framework version 1- Brief Synopsis and Tips &ndash; Part 1

    - by Rohit Gupta
    To Do Eager loading use Projections (for e.g. from c in context.Contacts select c, c.Addresses)  or use Include Query Builder Methods (Include(“Addresses”)) If there is multi-level hierarchical Data then to eager load all the relationships use Include Query Builder methods like customers.Include("Order.OrderDetail") to include Order and OrderDetail collections or use customers.Include("Order.OrderDetail.Location") to include all Order, OrderDetail and location collections with a single include statement =========================================================================== If the query uses Joins then Include() Query Builder method will be ignored, use Nested Queries instead If the query does projections then Include() Query Builder method will be ignored Use Address.ContactReference.Load() OR Contact.Addresses.Load() if you need to Deferred Load Specific Entity – This will result in extra round trips to the database ObjectQuery<> cannot return anonymous types... it will return a ObjectQuery<DBDataRecord> Only Include method can be added to Linq Query Methods Any Linq Query method can be added to Query Builder methods. If you need to append a Query Builder Method (other than Include) after a LINQ method  then cast the IQueryable<Contact> to ObjectQuery<Contact> and then append the Query Builder method to it =========================================================================== Query Builder methods are Select, Where, Include Methods which use Entity SQL as parameters e.g. "it.StartDate, it.EndDate" When Query Builder methods do projection then they return ObjectQuery<DBDataRecord>, thus to iterate over this collection use contact.Item[“Name”].ToString() When Linq To Entities methods do projection, they return collection of anonymous types --- thus the collection is strongly typed and supports Intellisense EF Object Context can track changes only on Entities, not on Anonymous types. If you use a Defining Query for a EntitySet then the EntitySet becomes readonly since a Defining Query is the same as a View (which is treated as a ReadOnly by default). However if you want to use this EntitySet for insert/update/deletes then we need to map stored procs (as created in the DB) to the insert/update/delete functions of the Entity in the Designer You can use either Execute method or ToList() method to bind data to datasources/bindingsources If you use the Execute Method then remember that you can traverse through the ObjectResult<> collection (returned by Execute) only ONCE. In WPF use ObservableCollection to bind to data sources , for keeping track of changes and letting EF send updates to the DB automatically. Use Extension Methods to add logic to Entities. For e.g. create extension methods for the EntityObject class. Create a method in ObjectContext Partial class and pass the entity as a parameter, then call this method as desired from within each entity. ================================================================ DefiningQueries and Stored Procedures: For Custom Entities, one can use DefiningQuery or Stored Procedures. Thus the Custom Entity Collection will be populated using the DefiningQuery (of the EntitySet) or the Sproc. If you use Sproc to populate the EntityCollection then the query execution is immediate and this execution happens on the Server side and any filters applied will be applied in the Client App. If we use a DefiningQuery then these queries are composable, meaning the filters (if applied to the entityset) will all be sent together as a single query to the DB, returning only filtered results. If the sproc returns results that cannot be mapped to existing entity, then we first create the Entity/EntitySet in the CSDL using Designer, then create a dummy Entity/EntitySet using XML in the SSDL. When creating a EntitySet in the SSDL for this dummy entity, use a TSQL that does not return any results, but does return the relevant columns e.g. select ContactID, FirstName, LastName from dbo.Contact where 1=2 Also insure that the Entity created in the SSDL uses the SQL DataTypes and not .NET DataTypes. If you are unable to open the EDMX file in the designer then note the Errors ... they will give precise info on what is wrong. The Thrid option is to simply create a Native Query in the SSDL using <Function Name="PaymentsforContact" IsComposable="false">   <CommandText>SELECT ActivityId, Activity AS ActivityName, ImagePath, Category FROM dbo.Activities </CommandText></FuncTion> Then map this Function to a existing Entity. This is a quick way to get a custom Entity which is regular Entity with renamed columns or additional columns (which are computed columns). The disadvantage to using this is that It will return all the rows from the Defining query and any filter (if defined) will be applied only at the Client side (after getting all the rows from DB). If you you DefiningQuery instead then we can use that as a Composable Query. The Fourth option (for mapping a READ stored proc results to a non-existent Entity) is to create a View in the Database which returns all the fields that the sproc also returns, then update the Model so that the model contains this View as a Entity. Then map the Read Sproc to this View Entity. The other option would be to simply create the View and remove the sproc altogether. ================================================================ To Execute a SProc that does not return a entity, use a EntityCommand to execute that proc. You cannot call a sproc FunctionImport that does not return Entities From Code, the only way is to use SSDL function calls using EntityCommand.  This changes with EntityFramework Version 4 where you can return Scalar Types, Complex Types, Entities or NonQuery ================================================================ UDF when created as a Function in SSDL, we need to set the Name & IsComposable properties for the Function element. IsComposable is always false for Sprocs, for UDF's set this to true. You cannot call UDF "Function" from within code since you cannot import a UDF Function into the CSDL Model (with Version 1 of EF). only stored procedures can be imported and then mapped to a entity ================================================================ Entity Framework requires properties that are involved in association mappings to be mapped in all of the function mappings for the entity (Insert, Update and Delete). Because Payment has an association to Reservation... hence we need to pass both the paymentId and reservationId to the Delete sproc even though just the paymentId is the PK on the Payment Table. ================================================================ When mapping insert, update and delete procs to a Entity, insure that all the three or none are mapped. Further if you have a base class and derived class in the CSDL, then you must map (ins, upd, del) sprocs to all parent and child entities in the inheritance relationship. Note that this limitation that base and derived entity methods must all must be mapped does not apply when you are mapping Read Stored Procedures.... ================================================================ You can write stored procedures SQL directly into the SSDL by creating a Function element in the SSDL and then once created, you can map this Function to a CSDL Entity directly in the designer during Function Import ================================================================ You can do Entity Splitting such that One Entity maps to multiple tables in the DB. For e.g. the Customer Entity currently derives from Contact Entity...in addition it also references the ContactPersonalInfo Entity. One can copy all properties from the ContactPersonalInfo Entity into the Customer Entity and then Delete the CustomerPersonalInfo entity, finall one needs to map the copied properties to the ContactPersonalInfo Table in Table Mapping (by adding another table (ContactPersonalInfo) to the Table Mapping... this is called Entity Splitting. Thus now when you insert a Customer record, it will automatically create SQL to insert records into the Contact, Customers and ContactPersonalInfo tables even though you have a Single Entity called Customer in the CSDL =================================================================== There is Table by Type Inheritance where another EDM Entity can derive from another EDM entity and absorb the inherted entities properties, for example in the Break Away Geek Adventures EDM, the Customer entity derives (inherits) from the Contact Entity and absorbs all the properties of Contact entity. Thus when you create a Customer Entity in Code and then call context.SaveChanges the Object Context will first create the TSQL to insert into the Contact Table followed by a TSQL to insert into the Customer table =================================================================== Then there is the Table per Hierarchy Inheritance..... where different types are created based on a condition (similar applying a condition to filter a Entity to contain filtered records)... the diference being that the filter condition populates a new Entity Type (derived from the base Entity). In the BreakAway sample the example is Lodging Entity which is a Abstract Entity and Then Resort and NonResort Entities which derive from Lodging Entity and records are filtered based on the value of the Resort Boolean field =================================================================== Then there is Table per Concrete Type Hierarchy where we create a concrete Entity for each table in the database. In the BreakAway sample there is a entity for the Reservation table and another Entity for the OldReservation table even though both the table contain the same number of fields. The OldReservation Entity can then inherit from the Reservation Entity and configure the OldReservation Entity to remove all Scalar Properties from the Entity (since it inherits the properties from Reservation and filters based on ReservationDate field) =================================================================== Complex Types (Complex Properties) Entities in EF can also contain Complex Properties (in addition to Scalar Properties) and these Complex Properties reference a ComplexType (not a EntityType) DropdownList, ListBox, RadioButtonList, CheckboxList, Bulletedlist are examples of List server controls (not data bound controls) these controls cannot use Complex properties during databinding, they need Scalar Properties. So if a Entity contains Complex properties and you need to bind those to list server controls then use projections to return Scalar properties and bind them to the control (the disadvantage is that projected collections are not tracked by the Object Context and hence cannot persist changes to the projected collections bound to controls) ObjectDataSource and EntityDataSource do account for Complex properties and one can bind entities with Complex Properties to Data Source controls and they will be tracked for changes... with no additional plumbing needed to persist changes to these collections bound to controls So DataBound controls like GridView, FormView need to use EntityDataSource or ObjectDataSource as a datasource for entities that contain Complex properties so that changes to the datasource done using the GridView can be persisted to the DB (enabling the controls for updates)....if you cannot use the EntityDataSource you need to flatten the ComplexType Properties using projections With EF Version 4 ComplexTypes are supported by the Designer and can add/remove/compose Complex Types directly using the Designer =================================================================== Conditional Mapping ... is like Table per Hierarchy Inheritance where Entities inherit from a base class and then used conditions to populate the EntitySet (called conditional Mapping). Conditional Mapping has limitations since you can only use =, is null and IS NOT NULL Conditions to do conditional mapping. If you need more operators for filtering/mapping conditionally then use QueryView(or possibly Defining Query) to create a readonly entity. QueryView are readonly by default... the EntitySet created by the QueryView is enabled for change tracking by the ObjectContext, however the ObjectContext cannot create insert/update/delete TSQL statements for these Entities when SaveChanges is called since it is QueryView. One way to get around this limitation is to map stored procedures for the insert/update/delete operations in the Designer. =================================================================== Difference between QueryView and Defining Query : QueryView is defined in the (MSL) Mapping File/section of the EDM XML, whereas the DefiningQuery is defined in the store schema (SSDL). QueryView is written using Entity SQL and is this database agnostic and can be used against any database/Data Layer. DefiningQuery is written using Database Lanaguage i.e. TSQL or PSQL thus you have more control =================================================================== Performance: Lazy loading is deferred loading done automatically. lazy loading is supported with EF version4 and is on by default. If you need to turn it off then use context.ContextOptions.lazyLoadingEnabled = false To improve Performance consider PreCompiling the ObjectQuery using the CompiledQuery.Compile method

    Read the article

  • How to Reuse Your Old Wi-Fi Router as a Network Switch

    - by Jason Fitzpatrick
    Just because your old Wi-Fi router has been replaced by a newer model doesn’t mean it needs to gather dust in the closet. Read on as we show you how to take an old and underpowered Wi-Fi router and turn it into a respectable network switch (saving your $20 in the process). Image by mmgallan. Why Do I Want To Do This? Wi-Fi technology has changed significantly in the last ten years but Ethernet-based networking has changed very little. As such, a Wi-Fi router with 2006-era guts is lagging significantly behind current Wi-Fi router technology, but the Ethernet networking component of the device is just as useful as ever; aside from potentially being only 100Mbs instead of 1000Mbs capable (which for 99% of home applications is irrelevant) Ethernet is Ethernet. What does this matter to you, the consumer? It means that even though your old router doesn’t hack it for your Wi-Fi needs any longer the device is still a perfectly serviceable (and high quality) network switch. When do you need a network switch? Any time you want to share an Ethernet cable among multiple devices, you need a switch. For example, let’s say you have a single Ethernet wall jack behind your entertainment center. Unfortunately you have four devices that you want to link to your local network via hardline including your smart HDTV, DVR, Xbox, and a little Raspberry Pi running XBMC. Instead of spending $20-30 to purchase a brand new switch of comparable build quality to your old Wi-Fi router it makes financial sense (and is environmentally friendly) to invest five minutes of your time tweaking the settings on the old router to turn it from a Wi-Fi access point and routing tool into a network switch–perfect for dropping behind your entertainment center so that your DVR, Xbox, and media center computer can all share an Ethernet connection. What Do I Need? For this tutorial you’ll need a few things, all of which you likely have readily on hand or are free for download. To follow the basic portion of the tutorial, you’ll need the following: 1 Wi-Fi router with Ethernet ports 1 Computer with Ethernet jack 1 Ethernet cable For the advanced tutorial you’ll need all of those things, plus: 1 copy of DD-WRT firmware for your Wi-Fi router We’re conducting the experiment with a Linksys WRT54GL Wi-Fi router. The WRT54 series is one of the best selling Wi-Fi router series of all time and there’s a good chance a significant number of readers have one (or more) of them stuffed in an office closet. Even if you don’t have one of the WRT54 series routers, however, the principles we’re outlining here apply to all Wi-Fi routers; as long as your router administration panel allows the necessary changes you can follow right along with us. A quick note on the difference between the basic and advanced versions of this tutorial before we proceed. Your typical Wi-Fi router has 5 Ethernet ports on the back: 1 labeled “Internet”, “WAN”, or a variation thereof and intended to be connected to your DSL/Cable modem, and 4 labeled 1-4 intended to connect Ethernet devices like computers, printers, and game consoles directly to the Wi-Fi router. When you convert a Wi-Fi router to a switch, in most situations, you’ll lose two port as the “Internet” port cannot be used as a normal switch port and one of the switch ports becomes the input port for the Ethernet cable linking the switch to the main network. This means, referencing the diagram above, you’d lose the WAN port and LAN port 1, but retain LAN ports 2, 3, and 4 for use. If you only need to switch for 2-3 devices this may be satisfactory. However, for those of you that would prefer a more traditional switch setup where there is a dedicated WAN port and the rest of the ports are accessible, you’ll need to flash a third-party router firmware like the powerful DD-WRT onto your device. Doing so opens up the router to a greater degree of modification and allows you to assign the previously reserved WAN port to the switch, thus opening up LAN ports 1-4. Even if you don’t intend to use that extra port, DD-WRT offers you so many more options that it’s worth the extra few steps. Preparing Your Router for Life as a Switch Before we jump right in to shutting down the Wi-Fi functionality and repurposing your device as a network switch, there are a few important prep steps to attend to. First, you want to reset the router (if you just flashed a new firmware to your router, skip this step). Following the reset procedures for your particular router or go with what is known as the “Peacock Method” wherein you hold down the reset button for thirty seconds, unplug the router and wait (while still holding the reset button) for thirty seconds, and then plug it in while, again, continuing to hold down the rest button. Over the life of a router there are a variety of changes made, big and small, so it’s best to wipe them all back to the factory default before repurposing the router as a switch. Second, after resetting, we need to change the IP address of the device on the local network to an address which does not directly conflict with the new router. The typical default IP address for a home router is 192.168.1.1; if you ever need to get back into the administration panel of the router-turned-switch to check on things or make changes it will be a real hassle if the IP address of the device conflicts with the new home router. The simplest way to deal with this is to assign an address close to the actual router address but outside the range of addresses that your router will assign via the DHCP client; a good pick then is 192.168.1.2. Once the router is reset (or re-flashed) and has been assigned a new IP address, it’s time to configure it as a switch. Basic Router to Switch Configuration If you don’t want to (or need to) flash new firmware onto your device to open up that extra port, this is the section of the tutorial for you: we’ll cover how to take a stock router, our previously mentioned WRT54 series Linksys, and convert it to a switch. Hook the Wi-Fi router up to the network via one of the LAN ports (consider the WAN port as good as dead from this point forward, unless you start using the router in its traditional function again or later flash a more advanced firmware to the device, the port is officially retired at this point). Open the administration control panel via  web browser on a connected computer. Before we get started two things: first,  anything we don’t explicitly instruct you to change should be left in the default factory-reset setting as you find it, and two, change the settings in the order we list them as some settings can’t be changed after certain features are disabled. To start, let’s navigate to Setup ->Basic Setup. Here you need to change the following things: Local IP Address: [different than the primary router, e.g. 192.168.1.2] Subnet Mask: [same as the primary router, e.g. 255.255.255.0] DHCP Server: Disable Save with the “Save Settings” button and then navigate to Setup -> Advanced Routing: Operating Mode: Router This particular setting is very counterintuitive. The “Operating Mode” toggle tells the device whether or not it should enable the Network Address Translation (NAT)  feature. Because we’re turning a smart piece of networking hardware into a relatively dumb one, we don’t need this feature so we switch from Gateway mode (NAT on) to Router mode (NAT off). Our next stop is Wireless -> Basic Wireless Settings: Wireless SSID Broadcast: Disable Wireless Network Mode: Disabled After disabling the wireless we’re going to, again, do something counterintuitive. Navigate to Wireless -> Wireless Security and set the following parameters: Security Mode: WPA2 Personal WPA Algorithms: TKIP+AES WPA Shared Key: [select some random string of letters, numbers, and symbols like JF#d$di!Hdgio890] Now you may be asking yourself, why on Earth are we setting a rather secure Wi-Fi configuration on a Wi-Fi router we’re not going to use as a Wi-Fi node? On the off chance that something strange happens after, say, a power outage when your router-turned-switch cycles on and off a bunch of times and the Wi-Fi functionality is activated we don’t want to be running the Wi-Fi node wide open and granting unfettered access to your network. While the chances of this are next-to-nonexistent, it takes only a few seconds to apply the security measure so there’s little reason not to. Save your changes and navigate to Security ->Firewall. Uncheck everything but Filter Multicast Firewall Protect: Disable At this point you can save your changes again, review the changes you’ve made to ensure they all stuck, and then deploy your “new” switch wherever it is needed. Advanced Router to Switch Configuration For the advanced configuration, you’ll need a copy of DD-WRT installed on your router. Although doing so is an extra few steps, it gives you a lot more control over the process and liberates an extra port on the device. Hook the Wi-Fi router up to the network via one of the LAN ports (later you can switch the cable to the WAN port). Open the administration control panel via web browser on the connected computer. Navigate to the Setup -> Basic Setup tab to get started. In the Basic Setup tab, ensure the following settings are adjusted. The setting changes are not optional and are required to turn the Wi-Fi router into a switch. WAN Connection Type: Disabled Local IP Address: [different than the primary router, e.g. 192.168.1.2] Subnet Mask: [same as the primary router, e.g. 255.255.255.0] DHCP Server: Disable In addition to disabling the DHCP server, also uncheck all the DNSMasq boxes as the bottom of the DHCP sub-menu. If you want to activate the extra port (and why wouldn’t you), in the WAN port section: Assign WAN Port to Switch [X] At this point the router has become a switch and you have access to the WAN port so the LAN ports are all free. Since we’re already in the control panel, however, we might as well flip a few optional toggles that further lock down the switch and prevent something odd from happening. The optional settings are arranged via the menu you find them in. Remember to save your settings with the save button before moving onto a new tab. While still in the Setup -> Basic Setup menu, change the following: Gateway/Local DNS : [IP address of primary router, e.g. 192.168.1.1] NTP Client : Disable The next step is to turn off the radio completely (which not only kills the Wi-Fi but actually powers the physical radio chip off). Navigate to Wireless -> Advanced Settings -> Radio Time Restrictions: Radio Scheduling: Enable Select “Always Off” There’s no need to create a potential security problem by leaving the Wi-Fi radio on, the above toggle turns it completely off. Under Services -> Services: DNSMasq : Disable ttraff Daemon : Disable Under the Security -> Firewall tab, uncheck every box except “Filter Multicast”, as seen in the screenshot above, and then disable SPI Firewall. Once you’re done here save and move on to the Administration tab. Under Administration -> Management:  Info Site Password Protection : Enable Info Site MAC Masking : Disable CRON : Disable 802.1x : Disable Routing : Disable After this final round of tweaks, save and then apply your settings. Your router has now been, strategically, dumbed down enough to plod along as a very dependable little switch. Time to stuff it behind your desk or entertainment center and streamline your cabling.     

    Read the article

  • Recreating OMS instances in a HA environment when instances on all nodes are lost

    - by rnigam
    Oracle highly recommends deploying EM in a HA environment. The best practices for HA deployments, backup and housekeeping of your Enterprise Manager environment are documented in the Oracle Enterprise Manager Advanced Configuration Guide. It is imperative that there is a good disaster recovery plan in place for your EM deployment. In this post I want to talk about a customer who failed to do the correct planning and housekeeping for EM and landed in a situation where we the all the OMSes were nearly blown away had we not jumped to help. We recently hit an issue at a customer site where we had a two node OMS setup of the Enterprise Manager and a RAC Database being used as the EM repository. An accidental delete of the OMS oracle home left us with a single node deployment. While we were trying to figure out a possible path to recover the first node, the second node was rebooted under a maintenance window. What followed was a complete site outage as the Admin and managed servers would not start on either of the nodes. In my situation there were - No backups of the Oracle Homes from any node - No OMS Configuration snapshots (created using the “emctl exportconfig oms” command) and the instance home was completely lost on node 1 which also had the Admin Server  We did however have: - A copy of the emkey.ora that I found under the OMS_ORACLE_HOME/ of the second node (NOTE: it is a bad practice to have your emkey present under the OMS Oracle home directory on the same server as the OMS. The backup of the emkey should be maintained on some other server. In this case however it was a savior in my situation since there were no backups - The oms oracle home on the second node but missing a number of files and had a number of changes done to the files in the home. There were a number of attempts to start the server by modifying various files based on the Weblogic server logs to have atleast node up and running but all of them failed. Here is how you can recover from this scenario: Follow these steps: STEP 1: Check status of emkey.ora Check whether the emkey exists is present in the EM repository or not. Run the following command: $OMS_ORACLE_HOME/bin/emctl status emkey If the output is something like this below then you are good to go and the key is present in the repository ./emctl status emkey Oracle Enterprise Manager 11g Release 1 Grid Control Copyright (c) 1996, 2010 Oracle Corporation. All rights reserved. Enter Enterprise Manager Root (SYSMAN) Password : The EMKey is configured properly. Here are the messages that you might see as the emctl status emkey output depending upon whether the EM Admin Server is up and if the key is configured properly: Case1:  AdminServer is up, emkey is proper in CredStore & not in repos. This is same as the output of the command shown above:The EMKey is configured properly Case 2: AdminServer is up, emkey is proper in CredStore & exists in repos:The EMKey is configured properly, but is not secure. Secure the EMKey by running "emctl config emkey -remove_from_repos".Case 3: AdminServer is down or emkey is corrupted in CredStore) & (emkey exists in repos): The EMKey exists in the Management Repository, but is not configured properly or is corrupted in the credential store.Configure the EMKey by running "emctl config emkey -copy_to_credstore".Case 4: (AdminServer is down or emkey is corrupted in CredStore) & (emkey does not exist in repos): The EMKey is not configured properly or is corrupted in the credential store and does not exist in the Management Repository. To correct the problem:1) Get the backed up emkey.ora file.2) Configure the emkey by running "emctl config emkey -copy_to_credstore_from_file". If not the key was not secured properly, we will have to be put in the repository before proceeding. Look at the next step 2 for doing this There may be cases (like mine) where running emctl may give errors like the following: $OMS_ORACLE_HOME/bin/emctl status emkey Exception in thread “Main Thread” java.lang.NoClassDefFoundError: oracle/security/pki/OracleWallet At oracle.sysman.emctl.config.oms.EMKeyCmds.main (EMKeyCmds.java:658) Just move to the next step to put the key back in the repository STEP 2: Put emkey.ora back in the repository Skip this step if your emkey.ora is present in the repository. If not, you need to put the key back in the repository See if you can run the following command (with sample output): $OMS_ORACLE_HOME/bin/emctl config emkey –copy_to_repos Oracle Enterprise Manager 11g Release 1 Grid Control Copyright (c) 1996, 2010 Oracle Corporation. All rights reserved. The EMKey has been copied to the Management Repository. This operation will cause the EMKey to become unsecure. After the required operation has been completed, secure the EMKey by running "emctl config emkey -remove_from_repos". Typically the key is present under $OMS_ORACLE_HOME/sysman/config directory before being removed after the install as a best practice. If you hit any errors while running emctl commands like the one mentioned in step 1, jump to step 3 and we will take care of the emkey.ora in Step 5 STEP 3: Get the port information Check for the existing port information in the emd.properties file under EM_INSTANCE_DIRECTORY (typically gc_inst directory right above the Middleware home where you have deployed em. For eg. /u01/app/oracle/product/gc_inst in case your oms home is /u01/app/oracle/product/Middleware/oms11g) In my case I got the information from the emgc.properties present in the gc_inst on the second node. If you can run emctl you may want to try the following command as well $OMS_ORACLE_HOME/bin/emctl status oms –details Note this information as this will be used in the next step STEP 4: Perform cleanup on Node 1 Note the oracle home of the Weblogic and OMS, get the list of applied patches in the homes (using opatch lsinventory command), take a backup copy of the home just in case we need it and then de-install/remove oracle homes, update inventory and cleanup processes on the first node STEP 5: Perform Software Only Installation of OMS on Node 1 Perform Weblogic 10.3.2 installation exactly under the same location as present in the earlier installation. Perform software only installation of the OMS using the following command. This will not run any configuration assistants and bypass all user interface validations runInstaller –noconfig -validationaswarnings Select the “Additional OMS” option while performing the installation. Provide the same path for OMS and Instance directories like the previous installation Use the port information collected in Step 3 while performing the installation. Once the installation is complete run the allroot.sh script to complete the binary deployment STEP 6: Apply one-off patches At this point you can apply any patches to the OMS Oracle Home previously. You only need to run opatch to install the patch in the home and not required to run the SQLs STEP 7: Copy EM key This step is only required if you were not able to use emctl command to put the emkey back into the EM repository in STEP 2 Copy the emkey.ora file of the old installation you have under $OMS_ORACLE_HOME/sysman/config directory of the newly installed OMS STEP 8: Configure Grid Control Domain Run the following command to configure the EM domain and OMS. Note that you need to use a different GC Domain name than what you used earlier. For example I have used GCDOMAIN11 as the new domain name when my previous domain name was GCDOMAIN $OMS_ORACLE_HOME/bin/omsca new –AS_USERNAME weblogic –EM_DOMAIN_NAME GCDOMAIN11 –NM_USER nodemanager -nostart This command as shown below will prompt for a number of inputs like Admin Server hostname, port, password, etc. Verify if the defaults shown are correct by pressing enter or provide a new value STEP 9: Run Add-ON Configuration Assistant After this step run the following add-on configuration assistant. This was used in my case to configure the virtualization add-on $OMS_ORACLE_HOME/addonca -oui -omsonly -name vt -install gc STEP 10: Start the OMS Now start the OMS using $OMS_ORACLE_HOME/bin/emctl start oms In a multi-node setup like mine you would either have a software load balancer or DNS round robin (using a virtual host name that resolves to one of multiple OMS hostnames) being used for load balancing. Secure the OMS against the SLB or DNS virtual hostname using the following $ OMS_HOME/bin/emctl secure oms -host slb.example.com -secure_port 1159 -slb_port 1159 -slb_console_port 443 STEP 11: Configure the Agent From the $AGENT_ORACLE_HOME/bin run the ./agentca –f At this point you should have your OMS on node 1 fully re-covered. Clean up node 2 and use the normal Additional OMS installation process documented in the official installation guide to add the additional OMS on node 2 Summary It took us nearly a little over two days to completely recover the environment with some other non-EM related issues that hit us along the way as well. In the end a situation like this could have been completely avoided had the proper housekeeping and backup of the Enterprise Manager Deployment been done in the first place. This is going to a topic that we cover in the next post. In the meantime please do refer to the Oracle Enterprise Manager Advanced Configuration Guide for planning your EM installation, backup and housekeeping procedures. This can be found here: http://download.oracle.com/docs/cd/E11857_01/index.htm Thanks This post would not have been possible without Raj Aggarwal, Prasad Chebrolu and Ravikumar Basa who helped to recover the environment and provided all the support we needed

    Read the article

< Previous Page | 63 64 65 66 67 68 69 70 71 72  | Next Page >