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Doubly linked lists

- by user1642677

I have an assignment that I am terribly lost on involving doubly linked lists (note, we are supposed to create it from scratch, not using built-in API's). The program is supposed to keep track of credit cards basically. My professor wants us to use doubly-linked lists to accomplish this. The problem is, the book does not go into detail on the subject (doesn't even show pseudo code involving doubly linked lists), it merely describes what a doubly linked list is and then talks with pictures and no code in a small paragraph. But anyway, I'm done complaining. I understand perfectly well how to create a node class and how it works. The problem is how do I use the nodes to create the list? Here is what I have so far. public class CardInfo { private String name; private String cardVendor; private String dateOpened; private double lastBalance; private int accountStatus; private final int MAX_NAME_LENGTH = 25; private final int MAX_VENDOR_LENGTH = 15; CardInfo() { } CardInfo(String n, String v, String d, double b, int s) { setName(n); setCardVendor(v); setDateOpened(d); setLastBalance(b); setAccountStatus(s); } public String getName() { return name; } public String getCardVendor() { return cardVendor; } public String getDateOpened() { return dateOpened; } public double getLastBalance() { return lastBalance; } public int getAccountStatus() { return accountStatus; } public void setName(String n) { if (n.length() > MAX_NAME_LENGTH) throw new IllegalArgumentException("Too Many Characters"); else name = n; } public void setCardVendor(String v) { if (v.length() > MAX_VENDOR_LENGTH) throw new IllegalArgumentException("Too Many Characters"); else cardVendor = v; } public void setDateOpened(String d) { dateOpened = d; } public void setLastBalance(double b) { lastBalance = b; } public void setAccountStatus(int s) { accountStatus = s; } public String toString() { return String.format("%-25s %-15s $%-s %-s %-s", name, cardVendor, lastBalance, dateOpened, accountStatus); } } public class CardInfoNode { CardInfo thisCard; CardInfoNode next; CardInfoNode prev; CardInfoNode() { } public void setCardInfo(CardInfo info) { thisCard.setName(info.getName()); thisCard.setCardVendor(info.getCardVendor()); thisCard.setLastBalance(info.getLastBalance()); thisCard.setDateOpened(info.getDateOpened()); thisCard.setAccountStatus(info.getAccountStatus()); } public CardInfo getInfo() { return thisCard; } public void setNext(CardInfoNode node) { next = node; } public void setPrev(CardInfoNode node) { prev = node; } public CardInfoNode getNext() { return next; } public CardInfoNode getPrev() { return prev; } } public class CardList { CardInfoNode head; CardInfoNode current; CardInfoNode tail; CardList() { head = current = tail = null; } public void insertCardInfo(CardInfo info) { if(head == null) { head = new CardInfoNode(); head.setCardInfo(info); head.setNext(tail); tail.setPrev(node) // here lies the problem. tail must be set to something // to make it doubly-linked. but tail is null since it's // and end point of the list. } } } Here is the assignment itself if it helps to clarify what is required and more importantly, the parts I'm not understanding. Thanks https://docs.google.com/open?id=0B3vVwsO0eQRaQlRSZG95eXlPcVE

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Keeping socket open to send files on timer calls?

- by user3704768

I'm writing a program that requires an image to be fetched from a remote server every 10 milliseconds or so, as that's how often the image is updated. My current method calls a timer to grab the image, but it encounters Socket Closed errors all the time, and sometimes does not work at all. How can I fix my methods to keep the socket open the whole time, so no reconnecting is needed? Here is the full class: import java.awt.event.ActionEvent; import java.awt.event.ActionListener; import java.io.File; import java.io.FileInputStream; import java.io.FileNotFoundException; import java.io.FileOutputStream; import java.io.IOException; import java.net.InetAddress; import java.net.InetSocketAddress; import java.net.ServerSocket; import java.net.Socket; import java.net.UnknownHostException; import javax.swing.Timer; public class Connection { public static void createServer() throws IOException { Capture.getScreen(); ServerSocket socket = null; try { socket = new ServerSocket(12345, 0, InetAddress.getByName("127.0.0.1")); } catch (UnknownHostException e) { e.printStackTrace(); } catch (IOException e) { e.printStackTrace(); } System.out.println("Server started on " + socket.getInetAddress().getHostAddress() + ":" + socket.getLocalPort() + ",\nWaiting for client to connect."); final Socket clientConnection = socket.accept(); System.out.println("Client accepted from " + clientConnection.getInetAddress().getHostAddress() + ", sending file"); ActionListener taskPerformer = new ActionListener() { public void actionPerformed(ActionEvent evt) { System.out.println("Sending File"); try { pipeStreams(new FileInputStream(new File( "captures/sCap.png")), clientConnection.getOutputStream(), 1024); } catch (FileNotFoundException e) { e.printStackTrace(); } catch (IOException e) { e.printStackTrace(); } } }; System.out.println("closing out connection"); try { clientConnection.close(); } catch (IOException e) { e.printStackTrace(); } try { socket.close(); } catch (IOException e) { e.printStackTrace(); } Timer timer = new Timer(10, taskPerformer); timer.setRepeats(true); timer.start(); } public static void createClient() throws IOException { System.out.println("Connecting to server."); final Socket socket = new Socket(); try { socket.connect(new InetSocketAddress(InetAddress .getByName("127.0.0.1"), 12345)); } catch (UnknownHostException e) { e.printStackTrace(); } catch (IOException e) { } ActionListener taskPerformer = new ActionListener() { public void actionPerformed(ActionEvent evt) { System.out.println("Success, retreiving file."); try { pipeStreams(socket.getInputStream(), new FileOutputStream( new File("captures/rCap.png")), 1024); } catch (FileNotFoundException e) { e.printStackTrace(); } catch (IOException e) { } } }; System.out.println("Closing connection"); try { socket.close(); } catch (IOException e) { e.printStackTrace(); } Timer timer = new Timer(10, taskPerformer); timer.setRepeats(true); timer.start(); } public static void pipeStreams(java.io.InputStream source, java.io.OutputStream destination, int bufferSize) throws IOException { byte[] buffer = new byte[bufferSize]; int read = 0; while ((read = source.read(buffer)) != -1) { destination.write(buffer, 0, read); } destination.flush(); destination.close(); source.close(); } }

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PHP & MySQL database storing the name Array problem.

- by comma

I'm new to PHP & MySQL and I keep getting the word Array stored in my MySQL tables fields $skill, $experience, $yearsI was wondering how can I fix this problem? And I know I need to add mysqli_real_escape_string which I left out to make code more easier to read Hopefully. Here is the PHP & MySQL code. if (isset($_POST['info_submitted'])) { $mysqli = mysqli_connect("localhost", "root", "", "sitename"); $dbc = mysqli_query($mysqli,"SELECT learned_skills.*, users_skills.* FROM learned_skills INNER JOIN users_skills ON learned_skills.id = users_skills.skill_id WHERE user_id='$user_id'"); if (!$dbc) { print mysqli_error($mysqli); return; } $user_id = '5'; $skill = $_POST['skill']; $experience = $_POST['experience']; $years = $_POST['years']; if (isset($_POST['skill'][0]) && trim($_POST['skill'][0])!=='') { $mysqli = mysqli_connect("localhost", "root", "", "sitename"); $dbc = mysqli_query($mysqli,"SELECT learned_skills.*, users_skills.* FROM learned_skills INNER JOIN users_skills ON users_skills.skill_id = learned_skills.id WHERE users_skills.user_id='$user_id'"); if (mysqli_num_rows($dbc) == 0) { for ($s = 0; $s < count($skill); $s++){ for ($x = 0; $x < count($experience); $x++){ for ($g = 0; $g < count($years); $g++){ $mysqli = mysqli_connect("localhost", "root", "", "sitename"); $query1 = mysqli_query($mysqli,"INSERT INTO learned_skills (skill, experience, years) VALUES ('" . $s . "', '" . $x . "', '" . $g . "')"); $id = mysqli_insert_id($mysqli); if ($query1 == TRUE) { $mysqli = mysqli_connect("localhost", "root", "", "sitename"); $query2 = mysqli_query($mysqli,"INSERT INTO users_skills (skill_id, user_id) VALUES ('$id', '$user_id')"); } } } } } } } Here is the XHTML code. <li><label for="skill">Skill: </label><input type="text" name="skill[0]" id="skill[0]" /> <label for="experience">Experience: </label> <?php echo '<select id="experience[0]" name="experience[0]">' . "\n"; foreach($experience_options as $option) { if ($option == $experience) { echo '<option value="' . $option . '" selected="selected">' . $option . '</option>' . "\n"; } else { echo '<option value="'. $option . '">' . $option . '</option>'."\n"; } } echo '</select>'; ?> <label for="years">Years: </label> <?php echo '<select id="years[0]" name="years[0]">' . "\n"; foreach($grade_options as $option) { if ($option == $years) { echo '<option value="' . $option . '" selected="selected">' . $option . '</option>' . "\n"; } else { echo '<option value="'. $option . '">' . $option . '</option>'."\n"; } } echo '</select>'; ?> </li>

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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

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load search results into a div that toggles with other divs

- by Z. Edwards

I am working with a page that has multiple divs that toggle. This function works. A search function was added and this works too. The problem with the page as it exists currently: The search bar was placed on the "default" div and the results load below the bar into another div that is invisible when empty. The results div is inside this first default div. If you toggle to another div, you lose the default div and can't get back to it. For this reason, I moved the search bar to the left navigation where the other toggle links are situated. I also moved the search results div out of the default div to "stand on its own." What I am trying to do: Make the search button show the div with the results as well as find the results. Basically, to integrate the search function into the array/toggle function. The search function is in one .js file and the toggle function is in a different .js file. I keep thinking there must be a way to get "onclick" to call from both .js files so that I don't have to do a bunch of extra work combining the two functions that already exist and work separately. I am a Javascript newbie learning by examples and haven't been able to figure this out. I have never seen a working example of this and my searches haven't produced one. I would be very grateful for any help. Hope I explained the problem adequately. Edit: Here is the code I already have for the toggle function. var ids=new Array('a','b','c',[and so on--search results not added here yet]); function switchid(id_array){ hideallids(); for( var i=0, limit=id_array.length; i < limit; ++i) showdiv(id_array[i]); } function hideallids(){ for (var i=0;i<ids.length;i++){ hidediv(ids[i]); } } function hidediv(id) { //safe function to hide an element with a specified id if (document.getElementById) { // DOM3 = IE5, NS6 document.getElementById(id).style.display = 'none'; } else { if (document.layers) { // Netscape 4 document.id.display = 'none'; } else { // IE 4 document.all.id.style.display = 'none'; } } } function showdiv(id) {//safe function to show an element with a specified id if (document.getElementById) { // DOM3 = IE5, NS6 document.getElementById(id).style.display = 'block'; } else { if (document.layers) { // Netscape 4 document.id.display = 'block'; } else { // IE 4 document.all.id.style.display = 'block'; } } } function initialize(){ var t = gup("target"); if( t ) { switchid([t]); } } function gup( name ) { name = name.replace(/[\[]/,"\\\[").replace(/[\]]/,"\\\]"); var regexS = "[\\?&]"+name+"=([^&#]*)"; var regex = new RegExp( regexS ); var results = regex.exec( window.location.href ); if( results == null ){ return ""; } else { return results[1]; } } Thanks in advance!

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Debugging errors in c++

- by user1513323

I was working on a program that printed out the word count, character count and line count depending on the user's input. But I keep getting these error that are completely unknown to me. I was wondering if anyone could help. ** I've changed it from previous mistakes and am still receiving errors. Sorry I'm new to C++. The errors I got were filestat.cpp:47: error: ‘line’ was not declared in this scope filestat.cpp: In function ‘int wc(std::string)’: filestat.cpp:55: error: ‘line’ was not declared in this scope filestat.cpp: In function ‘int cc(std::string)’: filestat.cpp:67: error: ‘line’ was not declared in this scope #include<iostream> #include<fstream> #include<string> using namespace std; int lc(string fname); int wc(string fname); int cc(string fname); int main(){ string fname,line,command; ifstream ifs; int i; while(true){ cout<<"---- Enter a file name : "; if(getline(cin,line)){ if(line.length()== 4 && line.compare("exit")== 0){ cout<<"Exiting"; exit(0); }else{ string command = line.substr(0,2); fname= line.substr(4, line.length() -5); if( ifs.fail()){ ifs.open(fname.c_str()); cerr<< "File not found" <<fname <<endl; ifs.clear(); }else{ if(command.compare("lc")){ lc(fname); }else if (command.compare("wc")){ wc(fname); }else if(command.compare("cc")){ cc(fname); }else cout<<"Command unknown. "; } } } } return 0; } int lc(string fname){ int count; while(getline(fname, line)){ count++; } cout<<"Number of lines: "<<count ; } int wc(string fname){ int count; while(getline(fname, line)){ int pos=line.find_first_of("\n\t ",0); while(pos =! string::npos){ int length=line.length(); line = line.substr(pos+1, length - pos); count++; } } cout<< "Number of words: " <<count; } int cc(string fname){ int count; while(getline(fname, line)){ count = count + line.length(); } cout<< "Number of words: " <<count; }

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Errors with parameter datatype in PostgreSql query

- by John

Im trying to execute a query to postgresql using the following code. It's written in C/C++ and I keep getting the following error when declaring a cursor: DECLARE CURSOR failed: ERROR: could not determine data type of parameter $1 Searching on here and on google, I can't find a solution. Can anyone find where I have made and error and why this is happening? thanks! void searchdb( PGconn *conn, char* name, char* offset ) { // Will hold the number of field in table int nFields; // Start a transaction block PGresult *res = PQexec(conn, "BEGIN"); if (PQresultStatus(res) != PGRES_COMMAND_OK) { printf("BEGIN command failed: %s", PQerrorMessage(conn)); PQclear(res); exit_nicely(conn); } // Clear result PQclear(res); printf("BEGIN command - OK\n"); //set the values to use const char *values[3] = {(char*)name, (char*)RESULTS_LIMIT, (char*)offset}; //calculate the lengths of each of the values int lengths[3] = {strlen((char*)name), sizeof(RESULTS_LIMIT), sizeof(offset)}; //state which parameters are binary int binary[3] = {0, 0, 1}; res = PQexecParams(conn, "DECLARE emprec CURSOR for SELECT name, id, 'Events' as source FROM events_basic WHERE name LIKE '$1::varchar%' UNION ALL " " SELECT name, fsq_id, 'Venues' as source FROM venues_cache WHERE name LIKE '$1::varchar%' UNION ALL " " SELECT name, geo_id, 'Cities' as source FROM static_cities WHERE name LIKE '$1::varchar%' OR FIND_IN_SET('$1::varchar%', alternate_names) != 0 LIMIT $2::int4 OFFSET $3::int4", 3, //number of parameters NULL, //ignore the Oid field values, //values to substitute $1 and $2 lengths, //the lengths, in bytes, of each of the parameter values binary, //whether the values are binary or not 0); //we want the result in text format // Fetch rows from table if (PQresultStatus(res) != PGRES_COMMAND_OK) { printf("DECLARE CURSOR failed: %s", PQerrorMessage(conn)); PQclear(res); exit_nicely(conn); } // Clear result PQclear(res); res = PQexec(conn, "FETCH ALL in emprec"); if (PQresultStatus(res) != PGRES_TUPLES_OK) { printf("FETCH ALL failed"); PQclear(res); exit_nicely(conn); } // Get the field name nFields = PQnfields(res); // Prepare the header with table field name printf("\nFetch record:"); printf("\n********************************************************************\n"); for (int i = 0; i < nFields; i++) printf("%-30s", PQfname(res, i)); printf("\n********************************************************************\n"); // Next, print out the record for each row for (int i = 0; i < PQntuples(res); i++) { for (int j = 0; j < nFields; j++) printf("%-30s", PQgetvalue(res, i, j)); printf("\n"); } PQclear(res); // Close the emprec res = PQexec(conn, "CLOSE emprec"); PQclear(res); // End the transaction res = PQexec(conn, "END"); // Clear result PQclear(res); }

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Select list shrinks in size horizontally when empty

- by joe

Hello, I have two select list boxes and i can move items back and forth between them using the forward (--) and backward (<--) button. However, if there are no items in a select list, it shrinks in size horizontally. Any way to keep the select list a fixed size, irrespective of whether it contains any options or not ? Below is the code: <html> <head> <script language="JavaScript" type="text/javascript">  </script> </head> <body> <form action="yourpage.asp" method="post"> <table border="0"> <tr> <td width="70"> <select name="sel1" size="10" multiple="multiple"> <option value="1">Left1</option> <option value="2">Left2</option> <option value="3">Left3</option> <option value="4">Left4</option> <option value="5">Left5</option> </select> </td> <td align="center" valign="middle"> <input type="button" value="-->" onclick="moveOptions(this.form.sel1, this.form.sel2);" /><br /> <input type="button" value="<--" onclick="moveOptions(this.form.sel2, this.form.sel1);" /> </td> <td> <select name="sel2" size="10" multiple="multiple"> <option value="1">Right1</option> <option value="2">Right2</option> <option value="3">Right3</option> <option value="4">Right4</option> <option value="5">Right5</option> </select> </td> </tr> </table> </form> </body> </html> Please help. Thank You.

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Incorrect data when passing pointer a list of pointers to a function. (C++)

- by Phil Elm

I'm writing code for combining data received over multiple sources. When the objects received (I'll call them MyPacket for now), they are stored in a standard list. However, whenever I reference the payload size of a partial MyPacket, the value shows up as 1 instead of the intended size. Here's the function code: MyPacket* CombinePackets(std::list<MyPacket*>* packets, uint8* current_packet){ uint32 total_payload_size = 0; if(packets->size() <= 0) return NULL; //For now. std::list<MyPacket*>::iterator it = packets->begin(); //Some minor code here, not relevant to the problem. for(uint8 index = 0; index < packets->size(); index++){ //(*it)->GetPayloadSize() returns 1 when it should show 1024. I've tried directly accessing the variable and more, but I just can't get it to work. total_payload_size += (*it)->GetPayloadSize(); cout << "Adding to total payload size value: " << (*it)->GetPayloadSize() << endl; std::advance(it,1); } MyPacket* packet = new MyPacket(); //Byte is just a typedef'd unsigned char. packet->payload = (byte) calloc(total_payload_size, sizeof(byte)); packet->payload_size = total_payload_size; it = packets->begin(); //Go back to the beginning again. uint32 big_payload_index = 0; for(uint8 index = 0; index < packets->size(); index++){ if(current_packet != NULL) *current_packet = index; for(uint32 payload_index = 0; payload_index < (*it)->GetPayloadSize(); payload_index++){ packet->payload[big_payload_index] = (*it)->payload[payload_index]; big_payload_index++; } std::advance(it,1); } return packet; } //Calling code std::list<MyPacket*> received = std::list<MyPacket*>(); //The code that fills it is here. std::list<MyPacket*>::iterator it = received.begin(); cout << (*it)->GetPayloadSize() << endl; // Outputs 1024 correctly! MyPacket* final = CombinePackets(&received,NULL); cout << final->GetPayloadSize() << endl; //Outputs 181, which happens to be the number of elements in the received list. So, as you can see above, when I reference (*it)-GetPayloadSize(), it returns 1 instead of the intended 1024. Can anyone see the problem and if so, do you have an idea on how to fix this? I've spent 4 hours searching and trying new solutions, but they all keep returning 1... EDIT:

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How to reduce iOS AVPlayer start delay

- by Bernt Habermeier

Note, for the below question: All assets are local on the device -- no network streaming is taking place. The videos contain audio tracks. I'm working on an iOS application that requires playing video files with minimum delay to start the video clip in question. Unfortunately we do not know what specific video clip is next until we actually need to start it up. Specifically: When one video clip is playing, we will know what the next set of (roughly) 10 video clips are, but we don't know which one exactly, until it comes time to 'immediately' play the next clip. What I've done to look at actual start delays is to call addBoundaryTimeObserverForTimes on the video player, with a time period of one millisecond to see when the video actually started to play, and I take the difference of that time stamp with the first place in the code that indicates which asset to start playing. From what I've seen thus-far, I have found that using the combination of AVAsset loading, and then creating an AVPlayerItem from that once it's ready, and then waiting for AVPlayerStatusReadyToPlay before I call play, tends to take between 1 and 3 seconds to start the clip. I've since switched to what I think is roughly equivalent: calling [AVPlayerItem playerItemWithURL:] and waiting for AVPlayerItemStatusReadyToPlay to play. Roughly same performance. One thing I'm observing is that the first AVPlayer item load is slower than the rest. Seems one idea is to pre-flight the AVPlayer with a short / empty asset before trying to play the first video might be of good general practice. [http://stackoverflow.com/questions/900461/slow-start-for-avaudioplayer-the-first-time-a-sound-is-played] I'd love to get the video start times down as much as possible, and have some ideas of things to experiment with, but would like some guidance from anyone that might be able to help. Update: idea 7, below, as-implemented yields switching times of around 500 ms. This is an improvement, but it it'd be nice to get this even faster. Idea 1: Use N AVPlayers (won't work) Using ~ 10 AVPPlayer objects and start-and-pause all ~ 10 clips, and once we know which one we really need, switch to, and un-pause the correct AVPlayer, and start all over again for the next cycle. I don't think this works, because I've read there is roughly a limit of 4 active AVPlayer's in iOS. There was someone asking about this on StackOverflow here, and found out about the 4 AVPlayer limit: fast-switching-between-videos-using-avfoundation Idea 2: Use AVQueuePlayer (won't work) I don't believe that shoving 10 AVPlayerItems into an AVQueuePlayer would pre-load them all for seamless start. AVQueuePlayer is a queue, and I think it really only makes the next video in the queue ready for immediate playback. I don't know which one out of ~10 videos we do want to play back, until it's time to start that one. ios-avplayer-video-preloading Idea 3: Load, Play, and retain AVPlayerItems in background (not 100% sure yet -- but not looking good) I'm looking at if there is any benefit to load and play the first second of each video clip in the background (suppress video and audio output), and keep a reference to each AVPlayerItem, and when we know which item needs to be played for real, swap that one in, and swap the background AVPlayer with the active one. Rinse and Repeat. The theory would be that recently played AVPlayer/AVPlayerItem's may still hold some prepared resources which would make subsequent playback faster. So far, I have not seen benefits from this, but I might not have the AVPlayerLayer setup correctly for the background. I doubt this will really improve things from what I've seen. Idea 4: Use a different file format -- maybe one that is faster to load? I'm currently using .m4v's (video-MPEG4) H.264 format. I have not played around with other formats, but it may well be that some formats are faster to decode / get ready than others. Possible still using video-MPEG4 but with a different codec, or maybe quicktime? Maybe a lossless video format where decoding / setup is faster? Idea 5: Combination of lossless video format + AVQueuePlayer If there is a video format that is fast to load, but maybe where the file size is insane, one idea might be to pre-prepare the first 10 seconds of each video clip with a version that is boated but faster to load, but back that up with an asset that is encoded in H.264. Use an AVQueuePlayer, and add the first 10 seconds in the uncompressed file format, and follow that up with one that is in H.264 which gets up to 10 seconds of prepare/preload time. So I'd get 'the best' of both worlds: fast start times, but also benefits from a more compact format. Idea 6: Use a non-standard AVPlayer / write my own / use someone else's Given my needs, maybe I can't use AVPlayer, but have to resort to AVAssetReader, and decode the first few seconds (possibly write raw file to disk), and when it comes to playback, make use of the raw format to play it back fast. Seems like a huge project to me, and if I go about it in a naive way, it's unclear / unlikely to even work better. Each decoded and uncompressed video frame is 2.25 MB. Naively speaking -- if we go with ~ 30 fps for the video, I'd end up with ~60 MB/s read-from-disk requirement, which is probably impossible / pushing it. Obviously we'd have to do some level of image compression (perhaps native openGL/es compression formats via PVRTC)... but that's kind crazy. Maybe there is a library out there that I can use? Idea 7: Combine everything into a single movie asset, and seekToTime One idea that might be easier than some of the above, is to combine everything into a single movie, and use seekToTime. The thing is that we'd be jumping all around the place. Essentially random access into the movie. I think this may actually work out okay: avplayer-movie-playing-lag-in-ios5 Which approach do you think would be best? So far, I've not made that much progress in terms of reducing the lag.

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Quick guide to Oracle IRM 11g: Configuring SSL

- by Simon Thorpe

Quick guide to Oracle IRM 11g index So far in this guide we have an IRM Server up and running, however I skipped over SSL configuration in the previous article because I wanted to focus in more detail now. You can, if you wish, not bother with setting up SSL, but considering this is a security technology it is worthwhile doing. Contents Setting up a one way, self signed SSL certificate in WebLogic Setting up an official SSL certificate in Apache 2.x Configuring Apache to proxy traffic to the IRM server There are two common scenarios in which an Oracle IRM server is configured. For a development or evaluation system, people usually communicate directly to the WebLogic Server running the IRM service. However in a production environment and for some proof of concept evaluations that require a setup reflecting a production system, the traffic to the IRM server travels via a web server proxy, commonly Apache. In this guide we are building an Oracle Enterprise Linux based IRM service and this article will go over the configuration of SSL in WebLogic and also in Apache. Like in the past articles, we are going to use two host names in the configuration below,irm.company.com will refer to the public Apache server irm.company.internal will refer to the internal WebLogic IRM server Setting up a one way, self signed SSL certificate in WebLogic First lets look at creating just a simple self signed SSL certificate to be used in WebLogic. This is a quick and easy way to get SSL working in your environment, however the downside is that no browsers are going to trust this certificate you create and you'll need to manually install the certificate onto any machine's communicating with the server. This is fine for development or when you have only a few users evaluating the system, but for any significant use it's usually better to have a fully trusted certificate in use and I explain that in the next section. But for now lets go through creating, installing and testing a self signed certificate. We use a library in Java to create the certificates, open a console and running the following commands. Note you should choose your own secure passwords whenever you see password below. [oracle@irm /] source /oracle/middleware/wlserver_10.3/server/bin/setWLSEnv.sh [oracle@irm /] cd /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig/ [oracle@irm /] java utils.CertGen -selfsigned -certfile MyOwnSelfCA.cer -keyfile MyOwnSelfKey.key -keyfilepass password -cn "irm.oracle.demo" [oracle@irm /] java utils.ImportPrivateKey -keystore MyOwnIdentityStore.jks -storepass password -keypass password -alias trustself -certfile MyOwnSelfCA.cer.pem -keyfile MyOwnSelfKey.key.pem -keyfilepass password [oracle@irm /] keytool -import -trustcacerts -alias trustself -keystore TrustMyOwnSelf.jks -file MyOwnSelfCA.cer.der -keyalg RSA We now have two Java Key Stores, MyOwnIdentityStore.jks and TrustMyOwnSelf.jks. These contain keys and certificates which we will use in WebLogic Server. Now we need to tell the IRM server to use these stores when setting up SSL connections for incoming requests. Make sure the Admin server is running and login into the WebLogic Console at http://irm.company.intranet:7001/console and do the following; In the menu on the left, select the + next to Environment to expose the submenu, then click on Servers. You will see two servers in the list, AdminServer(admin) and IRM_server1. If the IRM server is running, shut it down either by hitting CONTROL + C in the console window it was started from, or you can switch to the CONTROL tab, select IRM_server1 and then select the Shutdown menu and then Force Shutdown Now. In the Configuration tab select IRM_server1 and switch to the Keystores tab. By default WebLogic Server uses it's own demo identity and trust. We are now going to switch to the self signed one's we've just created. So select the Change button and switch to Custom Identity and Custom Trust and hit save. Now we have to complete the resulting fields, the setting's i've used in my evaluation server are below. IdentityCustom Identity Keystore: /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig/MyOwnIdentityStore.jks Custom Identity Keystore Type: JKS Custom Identity Keystore Passphrase: password Confirm Custom Identity Keystore Passphrase: password TrustCustom Trust Keystore: /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig/TrustMyOwnSelf.jks Custom Trust Keystore Type: JKS Custom Trust Keystore Passphrase: password Confirm Custom Trust Keystore Passphrase: password Now click on the SSL tab for the IRM_server1 and enter in the alias and passphrase, in my demo here the details are; IdentityPrivate Key Alias: trustself Private Key Passphrase: password Confirm Private Key Passphrase: password And hit save. Now lets test a connection to the IRM server over HTTPS using SSL. Go back to a console window and start the IRM server, a quick reminder on how to do this is... [oracle@irm /] cd /oracle/middleware/user_projects/domains/irm_domain/bin [oracle@irm /] ./startManagedWeblogic IRM_server1 Once running, open a browser and head to the SSL port of the server. By default the IRM server will be listening on the URL https://irm.company.intranet:16101/irm_rights. Note in the example image on the right the port is 7002 because it's a system that has the IRM services installed on the Admin server, this isn't typical (or advisable). Your system is going to have a separate managed server which will be listening on port 16101. Once you open this address you will notice that your browser is going to complain that the server certificate is untrusted. The images on the right show how Firefox displays this error. You are going to be prompted every time you create a new SSL session with the server, both from the browser and more annoyingly from the IRM Desktop. If you plan on always using a self signed certificate, it is worth adding it to the Windows certificate store so that when you are accessing sealed content you do not keep being informed this certificate is not trusted. Follow these instructions (which are for Internet Explorer 8, they may vary for your version of IE.) Start Internet Explorer and open the URL to your IRM server over SSL, e.g. https://irm.company.intranet:16101/irm_rights. IE will complain that about the certificate, click on Continue to this website (not recommended). From the IE Tools menu select Internet Options and from the resulting dialog select Security and then click on Trusted Sites and then the Sites button. Add to the list of trusted sites a URL which mates the server you are accessing, e.g. https://irm.company.intranet/ and select OK. Now refresh the page you were accessing and next to the URL you should see a red cross and the words Certificate Error. Click on this button and select View Certificates. You will now see a dialog with the details of the self signed certificate and the Install Certificate... button should be enabled. Click on this to start the wizard. Click next and you'll be asked where you should install the certificate. Change the option to Place all certificates in the following store. Select browse and choose the Trusted Root Certification Authorities location and hit OK. You'll then be prompted to install the certificate and answer yes. You also need to import the root signed certificate into the same location, so once again select the red Certificate Error option and this time when viewing the certificate, switch to the Certification Path tab and you should see a CertGenCAB certificate. Select this and then click on View Certificate and go through the same process as above to import the certificate into the store. Finally close all instances of the IE browser and re-access the IRM server URL again, this time you should not receive any errors. Setting up an official SSL certificate in Apache 2.x At this point we now have an IRM server that you can communicate with over SSL. However this certificate isn't trusted by any browser because it's path of trust doesn't end in a recognized certificate authority (CA). Also you are communicating directly to the WebLogic Server over a non standard SSL port, 16101. In a production environment it is common to have another device handle the initial public internet traffic and then proxy this to the WebLogic server. The diagram below shows a very simplified view of this type of deployment. What i'm going to walk through next is configuring Apache to proxy traffic to a WebLogic server and also to use a real SSL certificate from an official CA. First step is to configure Apache to handle incoming requests over SSL. In this guide I am configuring the IRM service in Oracle Enterprise Linux 5 update 3 and Apache 2.2.3 which came with OpenSSL and mod_ssl components. Before I purchase an SSL certificate, I need to generate a certificate request from the server. Oracle.com uses Verisign and for my own personal needs I use cheaper certificates from GoDaddy. The following instructions are specific to Apache, but there are many references out there for other web servers. For Apache I have OpenSSL and the commands are; [oracle@irm /] cd /usr/bin [oracle@irm bin] openssl genrsa -des3 -out irm-apache-server.key 2048 Generating RSA private key, 2048 bit long modulus ............................+++ .........+++ e is 65537 (0x10001) Enter pass phrase for irm-apache-server.key: Verifying - Enter pass phrase for irm-apache-server.key: [oracle@irm bin] openssl req -new -key irm-apache-server.key -out irm-apache-server.csr Enter pass phrase for irm-apache-server.key: You are about to be asked to enter information that will be incorporated into your certificate request. What you are about to enter is what is called a Distinguished Name or a DN. There are quite a few fields but you can leave some blank For some fields there will be a default value, If you enter '.', the field will be left blank. ----- Country Name (2 letter code) [GB]:US State or Province Name (full name) [Berkshire]:CA Locality Name (eg, city) [Newbury]:San Francisco Organization Name (eg, company) [My Company Ltd]:Oracle Organizational Unit Name (eg, section) []:Security Common Name (eg, your name or your server's hostname) []:irm.company.com Email Address []:[email protected] Please enter the following 'extra' attributes to be sent with your certificate request A challenge password []:testing An optional company name []: You must make sure to remember the pass phrase you used in the initial key generation, you will need this when later configuring Apache. In the /usr/bin directory there are now two new files. The irm-apache-server.csr contains our certificate request and is what you cut and paste, or upload, to your certificate authority when you purchase and validate your SSL certificate. In response you will typically get two files. Your server certificate and another certificate file that will likely contain a set of certificates from your CA which validate your certificate's trust. Next we need to configure Apache to use these files. Typically there is an ssl.conf file which is where all the SSL configuration is done. On my Oracle Enterprise Linux server this file is located in /etc/httpd/conf.d/ssl.conf and i've added the following lines. <VirtualHost irm.company.com> # Setup SSL for irm.company.com ServerName irm.company.com SSLEngine On SSLCertificateFile /oracle/secure/irm.company.com.crt SSLCertificateKeyFile /oracle/secure/irm.company.com.key SSLCertificateChainFile /oracle/secure/gd_bundle.crt </VirtualHost> Restarting Apache (apachectl restart) and I can now attempt to connect to the Apache server in a web browser, https://irm.company.com/. If all is configured correctly I should now see an Apache test page delivered to me over HTTPS. Configuring Apache to proxy traffic to the IRM server Final piece in setting up SSL is to have Apache proxy requests for the IRM server but do so securely. So the requests to Apache will be over HTTPS using a legitimate certificate, but we can also configure Apache to proxy these requests internally across to the IRM server using SSL with the self signed certificate we generated at the start of this article. To do this proxying we use the WebLogic Web Server plugin for Apache which you can download here from Oracle. Download the zip file and extract onto the server. The file extraction reveals a set of zip files, each one specific to a supported web server. In my instance I am using Apache 2.2 32bit on an Oracle Enterprise Linux, 64 bit server. If you are not sure what version your Apache server is, run the command /usr/sbin/httpd -V and you'll see version and it its 32 or 64 bit. Mine is a 32bit server so I need to extract the file WLSPlugin1.1-Apache2.2-linux32-x86.zip. The from the resulting lib folder copy the file mod_wl.so into /usr/lib/httpd/modules/. First we want to test that the plug in will work for regular HTTP traffic. Edit the httpd.conf for Apache and add the following section at the bottom. LoadModule weblogic_module modules/mod_wl.so <IfModule mod_weblogic.c> WebLogicHost irm.company.internal WebLogicPort 16100 WLLogFile /tmp/wl-proxy.log </IfModule> <Location /irm_rights> SetHandler weblogic-handler </Location> <Location /irm_desktop> SetHandler weblogic-handler </Location> <Location /irm_sealing> SetHandler weblogic-handler </Location> <Location /irm_services> SetHandler weblogic-handler </Location> Now restart Apache again (apachectl restart) and now open a browser to http://irm.company.com/irm_rights. Apache will proxy the HTTP traffic from the port 80 of your Apache server to the IRM service listening on port 16100 of the WebLogic Managed server. Note above I have included all four of the Locations you might wish to proxy. http://irm.company.internalirm_rights is the URL to the management website, /irm_desktop is the URL used for the IRM Desktop to communicate. irm_sealing is for web services based document sealing and irm_services is for IRM server web services. The last two are typically only used when you have the IRM server integrated with another application and it is unlikely you'd be accessing these resources from the public facing Apache server. However, just in case, i've mentioned them above. Now let's enable SSL communication from Apache to WebLogic. In the ZIP file we extracted were some more modules we need to copy into the Apache folder. Looking back in the lib that we extracted, there are some more files. Copy the following into the /usr/lib/httpd/modules/ folder. libwlssl.so libnnz11.so libclntsh.so.11.1 Now the documentation states that should only need to do this, but I found that I also needed to create an environment variable called LD_LIBRARY_PATH and point this to the folder /usr/lib/httpd/modules/. If I didn't do this, starting Apache with the WebLogic module configured to SSL would throw the error. [crit] (20014)Internal error: WL SSL Init failed for server: (null) on 0 So I had to edit the file /etc/profile and add the following lines at the bottom. You may already have the LD_LIBRARY_PATH variable defined, therefore simply add this path to it. LD_LIBRARY_PATH=/usr/lib/httpd/modules/ export LD_LIBRARY_PATH Now the WebLogic plug in uses an Oracle Wallet to store the required certificates.You'll need to copy the self signed certificate from the IRM server over to the Apache server. Copy over the MyOwnSelfCA.cer.der into the same folder where you are storing your public certificates, in my example this is /oracle/secure. It's worth mentioning these files should ONLY be readable by root (the user Apache runs as). Now lets create an Oracle Wallet and import the self signed certificate from the IRM server. The file orapki was included in the bin folder of the Apache 1.1 plugin zip you extracted. orapki wallet create -wallet /oracle/secure/my-wallet -auto_login_only orapki wallet add -wallet /oracle/secure/my-wallet -trusted_cert -cert MyOwnSelfCA.cer.der -auto_login_only Finally change the httpd.conf to reflect that we want the WebLogic Apache plug-in to use HTTPS/SSL and not just plain HTTP. <IfModule mod_weblogic.c> WebLogicHost irm.company.internal WebLogicPort 16101 SecureProxy ON WLSSLWallet /oracle/secure/my-wallet WLLogFile /tmp/wl-proxy.log </IfModule> Then restart Apache once more and you can go back to the browser to test the communication. Opening the URL https://irm.company.com/irm_rights will proxy your request to the WebLogic server at https://irm.company.internal:16101/irm_rights. At this point you have a fully functional Oracle IRM service, the next step is to create a sealed document and test the entire system.

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HTG Reviews the CODE Keyboard: Old School Construction Meets Modern Amenities

- by Jason Fitzpatrick

There’s nothing quite as satisfying as the smooth and crisp action of a well built keyboard. If you’re tired of mushy keys and cheap feeling keyboards, a well-constructed mechanical keyboard is a welcome respite from the $10 keyboard that came with your computer. Read on as we put the CODE mechanical keyboard through the paces. What is the CODE Keyboard? The CODE keyboard is a collaboration between manufacturer WASD Keyboards and Jeff Atwood of Coding Horror (the guy behind the Stack Exchange network and Discourse forum software). Atwood’s focus was incorporating the best of traditional mechanical keyboards and the best of modern keyboard usability improvements. In his own words: The world is awash in terrible, crappy, no name how-cheap-can-we-make-it keyboards. There are a few dozen better mechanical keyboard options out there. I’ve owned and used at least six different expensive mechanical keyboards, but I wasn’t satisfied with any of them, either: they didn’t have backlighting, were ugly, had terrible design, or were missing basic functions like media keys. That’s why I originally contacted Weyman Kwong of WASD Keyboards way back in early 2012. I told him that the state of keyboards was unacceptable to me as a geek, and I proposed a partnership wherein I was willing to work with him to do whatever it takes to produce a truly great mechanical keyboard. Even the ardent skeptic who questions whether Atwood has indeed created a truly great mechanical keyboard certainly can’t argue with the position he starts from: there are so many agonizingly crappy keyboards out there. Even worse, in our opinion, is that unless you’re a typist of a certain vintage there’s a good chance you’ve never actually typed on a really nice keyboard. Those that didn’t start using computers until the mid-to-late 1990s most likely have always typed on modern mushy-key keyboards and never known the joy of typing on a really responsive and crisp mechanical keyboard. Is our preference for and love of mechanical keyboards shining through here? Good. We’re not even going to try and hide it. So where does the CODE keyboard stack up in pantheon of keyboards? Read on as we walk you through the simple setup and our experience using the CODE. Setting Up the CODE Keyboard Although the setup of the CODE keyboard is essentially plug and play, there are two distinct setup steps that you likely haven’t had to perform on a previous keyboard. Both highlight the degree of care put into the keyboard and the amount of customization available. Inside the box you’ll find the keyboard, a micro USB cable, a USB-to-PS2 adapter, and a tool which you may be unfamiliar with: a key puller. We’ll return to the key puller in a moment. Unlike the majority of keyboards on the market, the cord isn’t permanently affixed to the keyboard. What does this mean for you? Aside from the obvious need to plug it in yourself, it makes it dead simple to repair your own keyboard cord if it gets attacked by a pet, mangled in a mechanism on your desk, or otherwise damaged. It also makes it easy to take advantage of the cable routing channels in on the underside of the keyboard to route your cable exactly where you want it. While we’re staring at the underside of the keyboard, check out those beefy rubber feet. By peripherals standards they’re huge (and there is six instead of the usual four). Once you plunk the keyboard down where you want it, it might as well be glued down the rubber feet work so well. After you’ve secured the cable and adjusted it to your liking, there is one more task before plug the keyboard into the computer. On the bottom left-hand side of the keyboard, you’ll find a small recess in the plastic with some dip switches inside: The dip switches are there to switch hardware functions for various operating systems, keyboard layouts, and to enable/disable function keys. By toggling the dip switches you can change the keyboard from QWERTY mode to Dvorak mode and Colemak mode, the two most popular alternative keyboard configurations. You can also use the switches to enable Mac-functionality (for Command/Option keys). One of our favorite little toggles is the SW3 dip switch: you can disable the Caps Lock key; goodbye accidentally pressing Caps when you mean to press Shift. You can review the entire dip switch configuration chart here. The quick-start for Windows users is simple: double check that all the switches are in the off position (as seen in the photo above) and then simply toggle SW6 on to enable the media and backlighting function keys (this turns the menu key on the keyboard into a function key as typically found on laptop keyboards). After adjusting the dip switches to your liking, plug the keyboard into an open USB port on your computer (or into your PS/2 port using the included adapter). Design, Layout, and Backlighting The CODE keyboard comes in two flavors, a traditional 87-key layout (no number pad) and a traditional 104-key layout (number pad on the right hand side). We identify the layout as traditional because, despite some modern trapping and sneaky shortcuts, the actual form factor of the keyboard from the shape of the keys to the spacing and position is as classic as it comes. You won’t have to learn a new keyboard layout and spend weeks conditioning yourself to a smaller than normal backspace key or a PgUp/PgDn pair in an unconventional location. Just because the keyboard is very conventional in layout, however, doesn’t mean you’ll be missing modern amenities like media-control keys. The following additional functions are hidden in the F11, F12, Pause button, and the 2×6 grid formed by the Insert and Delete rows: keyboard illumination brightness, keyboard illumination on/off, mute, and then the typical play/pause, forward/backward, stop, and volume +/- in Insert and Delete rows, respectively. While we weren’t sure what we’d think of the function-key system at first (especially after retiring a Microsoft Sidewinder keyboard with a huge and easily accessible volume knob on it), it took less than a day for us to adapt to using the Fn key, located next to the right Ctrl key, to adjust our media playback on the fly. Keyboard backlighting is a largely hit-or-miss undertaking but the CODE keyboard nails it. Not only does it have pleasant and easily adjustable through-the-keys lighting but the key switches the keys themselves are attached to are mounted to a steel plate with white paint. Enough of the light reflects off the interior cavity of the keys and then diffuses across the white plate to provide nice even illumination in between the keys. Highlighting the steel plate beneath the keys brings us to the actual construction of the keyboard. It’s rock solid. The 87-key model, the one we tested, is 2.0 pounds. The 104-key is nearly a half pound heavier at 2.42 pounds. Between the steel plate, the extra-thick PCB board beneath the steel plate, and the thick ABS plastic housing, the keyboard has very solid feel to it. Combine that heft with the previously mentioned thick rubber feet and you have a tank-like keyboard that won’t budge a millimeter during normal use. Examining The Keys This is the section of the review the hardcore typists and keyboard ninjas have been waiting for. We’ve looked at the layout of the keyboard, we’ve looked at the general construction of it, but what about the actual keys? There are a wide variety of keyboard construction techniques but the vast majority of modern keyboards use a rubber-dome construction. The key is floated in a plastic frame over a rubber membrane that has a little rubber dome for each key. The press of the physical key compresses the rubber dome downwards and a little bit of conductive material on the inside of the dome’s apex connects with the circuit board. Despite the near ubiquity of the design, many people dislike it. The principal complaint is that dome keyboards require a complete compression to register a keystroke; keyboard designers and enthusiasts refer to this as “bottoming out”. In other words, the register the “b” key, you need to completely press that key down. As such it slows you down and requires additional pressure and movement that, over the course of tens of thousands of keystrokes, adds up to a whole lot of wasted time and fatigue. The CODE keyboard features key switches manufactured by Cherry, a company that has manufactured key switches since the 1960s. Specifically the CODE features Cherry MX Clear switches. These switches feature the same classic design of the other Cherry switches (such as the MX Blue and Brown switch lineups) but they are significantly quieter (yes this is a mechanical keyboard, but no, your neighbors won’t think you’re firing off a machine gun) as they lack the audible click found in most Cherry switches. This isn’t to say that they keyboard doesn’t have a nice audible key press sound when the key is fully depressed, but that the key mechanism isn’t doesn’t create a loud click sound when triggered. One of the great features of the Cherry MX clear is a tactile “bump” that indicates the key has been compressed enough to register the stroke. For touch typists the very subtle tactile feedback is a great indicator that you can move on to the next stroke and provides a welcome speed boost. Even if you’re not trying to break any word-per-minute records, that little bump when pressing the key is satisfying. The Cherry key switches, in addition to providing a much more pleasant typing experience, are also significantly more durable than dome-style key switch. Rubber dome switch membrane keyboards are typically rated for 5-10 million contacts whereas the Cherry mechanical switches are rated for 50 million contacts. You’d have to write the next War and Peace and follow that up with A Tale of Two Cities: Zombie Edition, and then turn around and transcribe them both into a dozen different languages to even begin putting a tiny dent in the lifecycle of this keyboard. So what do the switches look like under the classicly styled keys? You can take a look yourself with the included key puller. Slide the loop between the keys and then gently beneath the key you wish to remove: Wiggle the key puller gently back and forth while exerting a gentle upward pressure to pop the key off; You can repeat the process for every key, if you ever find yourself needing to extract piles of cat hair, Cheeto dust, or other foreign objects from your keyboard. There it is, the naked switch, the source of that wonderful crisp action with the tactile bump on each keystroke. The last feature worthy of a mention is the N-key rollover functionality of the keyboard. This is a feature you simply won’t find on non-mechanical keyboards and even gaming keyboards typically only have any sort of key roller on the high-frequency keys like WASD. So what is N-key rollover and why do you care? On a typical mass-produced rubber-dome keyboard you cannot simultaneously press more than two keys as the third one doesn’t register. PS/2 keyboards allow for unlimited rollover (in other words you can’t out type the keyboard as all of your keystrokes, no matter how fast, will register); if you use the CODE keyboard with the PS/2 adapter you gain this ability. If you don’t use the PS/2 adapter and use the native USB, you still get 6-key rollover (and the CTRL, ALT, and SHIFT don’t count towards the 6) so realistically you still won’t be able to out type the computer as even the more finger twisting keyboard combos and high speed typing will still fall well within the 6-key rollover. The rollover absolutely doesn’t matter if you’re a slow hunt-and-peck typist, but if you’ve read this far into a keyboard review there’s a good chance that you’re a serious typist and that kind of quality construction and high-number key rollover is a fantastic feature. The Good, The Bad, and the Verdict We’ve put the CODE keyboard through the paces, we’ve played games with it, typed articles with it, left lengthy comments on Reddit, and otherwise used and abused it like we would any other keyboard. The Good: The construction is rock solid. In an emergency, we’re confident we could use the keyboard as a blunt weapon (and then resume using it later in the day with no ill effect on the keyboard). The Cherry switches are an absolute pleasure to type on; the Clear variety found in the CODE keyboard offer a really nice middle-ground between the gun-shot clack of a louder mechanical switch and the quietness of a lesser-quality dome keyboard without sacrificing quality. Touch typists will love the subtle tactile bump feedback. Dip switch system makes it very easy for users on different systems and with different keyboard layout needs to switch between operating system and keyboard layouts. If you’re investing a chunk of change in a keyboard it’s nice to know you can take it with you to a different operating system or “upgrade” it to a new layout if you decide to take up Dvorak-style typing. The backlighting is perfect. You can adjust it from a barely-visible glow to a blazing light-up-the-room brightness. Whatever your intesity preference, the white-coated steel backplate does a great job diffusing the light between the keys. You can easily remove the keys for cleaning (or to rearrange the letters to support a new keyboard layout). The weight of the unit combined with the extra thick rubber feet keep it planted exactly where you place it on the desk. The Bad: While you’re getting your money’s worth, the $150 price tag is a shock when compared to the $20-60 price tags you find on lower-end keyboards. People used to large dedicated media keys independent of the traditional key layout (such as the large buttons and volume controls found on many modern keyboards) might be off put by the Fn-key style media controls on the CODE. The Verdict: The keyboard is clearly and heavily influenced by the needs of serious typists. Whether you’re a programmer, transcriptionist, or just somebody that wants to leave the lengthiest article comments the Internet has ever seen, the CODE keyboard offers a rock solid typing experience. Yes, $150 isn’t pocket change, but the quality of the CODE keyboard is so high and the typing experience is so enjoyable, you’re easily getting ten times the value you’d get out of purchasing a lesser keyboard. Even compared to other mechanical keyboards on the market, like the Das Keyboard, you’re still getting more for your money as other mechanical keyboards don’t come with the lovely-to-type-on Cherry MX Clear switches, back lighting, and hardware-based operating system keyboard layout switching. If it’s in your budget to upgrade your keyboard (especially if you’ve been slogging along with a low-end rubber-dome keyboard) there’s no good reason to not pickup a CODE keyboard. Key animation courtesy of Geekhack.org user Lethal Squirrel.

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Maintain cookie session in Android

- by datguywhowanders

Okay, I have an android application that has a form in it, two EditText, a spinner, and a login button. The user selects the service from the spinner, types in their user name and password, and clicks login. The data is sent via POST, a response is returned, it's handled, a new webview is launched, the html string generated form the response is loaded, and I have the home page of whatever service the user selected. That's all well and good. Now, when the user clicks on a link, the login info can't be found, and the page asks the user to login again. My login session is being dropped somewhere, and I'm not certain how to pass the info from the class that controls the main part of my app to the class that just launches the webview activity. The on click handler from the form login button: private class FormOnClickListener implements View.OnClickListener { public void onClick(View v) { String actionURL, user, pwd, user_field, pwd_field; actionURL = "thePageURL"; user_field = "username"; //this changes based on selections in a spinner pwd_field = "password"; //this changes based on selections in a spinner user = "theUserLogin"; pwd = "theUserPassword"; List<NameValuePair> myList = new ArrayList<NameValuePair>(); myList.add(new BasicNameValuePair(user_field, user)); myList.add(new BasicNameValuePair(pwd_field, pwd)); HttpParams params = new BasicHttpParams(); DefaultHttpClient client = new DefaultHttpClient(params); HttpPost post = new HttpPost(actionURL); HttpResponse response = null; BasicResponseHandler myHandler = new BasicResponseHandler(); String endResult = null; try { post.setEntity(new UrlEncodedFormEntity(myList)); } catch (UnsupportedEncodingException e) { e.printStackTrace(); } try { response = client.execute(post); } catch (ClientProtocolException e) { e.printStackTrace(); } catch (IOException e) { e.printStackTrace(); } try { endResult = myHandler.handleResponse(response); } catch (HttpResponseException e) { e.printStackTrace(); } catch (IOException e) { e.printStackTrace(); } List cookies = client.getCookieStore().getCookies(); if (!cookies.isEmpty()) { for (int i = 0; i < cookies.size(); i++) { cookie = cookies.get(i); } } Intent myWebViewIntent = new Intent(MsidePortal.this, MyWebView.class); myWebViewIntent.putExtra("htmlString", endResult); myWebViewIntent.putExtra("actionURL", actionURL); startActivity(myWebViewIntent); } } And here is the webview class that handles the response display: public class MyWebView extends android.app.Activity{ private class MyWebViewClient extends WebViewClient { @Override public boolean shouldOverrideUrlLoading(WebView view, String url) { view.loadUrl(url); return true; } } @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.web); MyWebViewClient myClient = new MyWebViewClient(); WebView webview = (WebView)findViewById(R.id.mainwebview); webview.getSettings().setBuiltInZoomControls(true); webview.getSettings().setJavaScriptEnabled(true); webview.setWebViewClient(myClient); Bundle extras = getIntent().getExtras(); if(extras != null) { // Get endResult String htmlString = extras.getString("htmlString"); String actionURL = extras.getString("actionURL"); Cookie sessionCookie = MsidePortal.cookie; CookieSyncManager.createInstance(this); CookieManager cookieManager = CookieManager.getInstance(); if (sessionCookie != null) { cookieManager.removeSessionCookie(); String cookieString = sessionCookie.getName() + "=" + sessionCookie.getValue() + "; domain=" + sessionCookie.getDomain(); cookieManager.setCookie(actionURL, cookieString); CookieSyncManager.getInstance().sync(); } webview.loadDataWithBaseURL(actionURL, htmlString, "text/html", "utf-8", actionURL); } } } I've had mixed success implementing that cookie solution. It seems to work for one service I log into that I know keeps the cookies on the server (old, archaic, but it works and they don't want to change it.) The service I'm attempting now requires the user to keep cookies on their local machine, and it does not work with this setup. Any suggestions?

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Setting up and using Bing Translate API Service for Machine Translation

- by Rick Strahl

Last week I spent quite a bit of time trying to set up the Bing Translate API service. I can honestly say this was one of the most screwed up developer experiences I've had in a long while - specifically related to the byzantine sign up process that Microsoft has in place. Not only is it nearly impossible to find decent documentation on the required signup process, some of the links in the docs are just plain wrong, and some of the account pages you need to access the actual account information once signed up are not linked anywhere from the administration UI. To make things even harder is the fact that the APIs changed a while back, with a completely new authentication scheme that's described and not directly linked documentation topic also made for a very frustrating search experience. It's a bummer that this is the case too, because the actual API itself is easy to use and works very well - fast and reasonably accurate (as accurate as you can expect machine translation to be). But the sign up process is a pain in the ass doubtlessly leaving many people giving up in frustration. In this post I'll try to hit all the points needed to set up to use the Bing Translate API in one place since such a document seems to be missing from Microsoft. Hopefully the API folks at Microsoft will get their shit together and actually provide this sort of info on their site… Signing Up The first step required is to create a Windows Azure MarketPlace account. Go to: https://datamarket.azure.com/ Sign in with your Windows Live Id If you don't have an account you will be taken to a registration page which you have to fill out. Follow the links and complete the registration. Once you're signed in you can start adding services. Click on the Data Link on the main page Select Microsoft Translator from the list This adds the Microsoft Bing Translator to your services. Pricing The page shows the pricing matrix and the free service which provides 2 megabytes for translations a month for free. Prices go up steeply from there. Pricing is determined by actual bytes of the result translations used. Max translations are 1000 characters so at minimum this means you get around 2000 translations a month for free. However most translations are probable much less so you can expect larger number of translations to go through. For testing or low volume translations this should be just fine. Once signed up there are no further instructions and you're left in limbo on the MS site. Register your Application Once you've created the Data association with Translator the next step is registering your application. To do this you need to access your developer account. Go to https://datamarket.azure.com/developer/applications/register Provide a ClientId, which is effectively the unique string identifier for your application (not your customer id!) Provide your name The client secret was auto-created and this becomes your 'password' For the redirect url provide any https url: https://microsoft.com works Give this application a description of your choice so you can identify it in the list of apps Now, once you've registered your application, keep track of the ClientId and ClientSecret - those are the two keys you need to authenticate before you can call the Translate API. Oddly the applications page is hidden from the Azure Portal UI. I couldn't find a direct link from anywhere on the site back to this page where I can examine my developer application keys. To find them you can go to: https://datamarket.azure.com/developer/applications You can come back here to look at your registered applications and pick up the ClientID and ClientSecret. Fun eh? But we're now ready to actually call the API and do some translating. Using the Bing Translate API The good news is that after this signup hell, using the API is pretty straightforward. To use the translation API you'll need to actually use two services: You need to call an authentication API service first, before you can call the actual translator API. These two APIs live on different domains, and the authentication API returns JSON data while the translator service returns XML. So much for consistency. Authentication The first step is authentication. The service uses oAuth authentication with a bearer token that has to be passed to the translator API. The authentication call retrieves the oAuth token that you can then use with the translate API call. The bearer token has a short 10 minute life time, so while you can cache it for successive calls, the token can't be cached for long periods. This means for Web backend requests you typically will have to authenticate each time unless you build a more elaborate caching scheme that takes the timeout into account (perhaps using the ASP.NET Cache object). For low volume operations you can probably get away with simply calling the auth API for every translation you do. To call the Authentication API use code like this:/// /// Retrieves an oAuth authentication token to be used on the translate /// API request. The result string needs to be passed as a bearer token /// to the translate API. /// /// You can find client ID and Secret (or register a new one) at: /// https://datamarket.azure.com/developer/applications/ /// /// The client ID of your application /// The client secret or password /// public string GetBingAuthToken(string clientId = null, string clientSecret = null) { string authBaseUrl = https://datamarket.accesscontrol.windows.net/v2/OAuth2-13; if (string.IsNullOrEmpty(clientId) || string.IsNullOrEmpty(clientSecret)) { ErrorMessage = Resources.Resources.Client_Id_and_Client_Secret_must_be_provided; return null; } var postData = string.Format("grant_type=client_credentials&client_id={0}" + "&client_secret={1}" + "&scope=http://api.microsofttranslator.com", HttpUtility.UrlEncode(clientId), HttpUtility.UrlEncode(clientSecret)); // POST Auth data to the oauth API string res, token; try { var web = new WebClient(); web.Encoding = Encoding.UTF8; res = web.UploadString(authBaseUrl, postData); } catch (Exception ex) { ErrorMessage = ex.GetBaseException().Message; return null; } var ser = new JavaScriptSerializer(); var auth = ser.Deserialize<BingAuth>(res); if (auth == null) return null; token = auth.access_token; return token; } private class BingAuth { public string token_type { get; set; } public string access_token { get; set; } } This code basically takes the client id and secret and posts it at the oAuth endpoint which returns a JSON string. Here I use the JavaScript serializer to deserialize the JSON into a custom object I created just for deserialization. You can also use JSON.NET and dynamic deserialization if you are already using JSON.NET in your app in which case you don't need the extra type. In my library that houses this component I don't, so I just rely on the built in serializer. The auth method returns a long base64 encoded string which can be used as a bearer token in the translate API call. Translation Once you have the authentication token you can use it to pass to the translate API. The auth token is passed as an Authorization header and the value is prefixed with a 'Bearer ' prefix for the string. Here's what the simple Translate API call looks like:/// /// Uses the Bing API service to perform translation /// Bing can translate up to 1000 characters. /// /// Requires that you provide a CLientId and ClientSecret /// or set the configuration values for these two. /// /// More info on setup: /// http://www.west-wind.com/weblog/ /// /// Text to translate /// Two letter culture name /// Two letter culture name /// Pass an access token retrieved with GetBingAuthToken. /// If not passed the default keys from .config file are used if any /// public string TranslateBing(string text, string fromCulture, string toCulture, string accessToken = null) { string serviceUrl = "http://api.microsofttranslator.com/V2/Http.svc/Translate"; if (accessToken == null) { accessToken = GetBingAuthToken(); if (accessToken == null) return null; } string res; try { var web = new WebClient(); web.Headers.Add("Authorization", "Bearer " + accessToken); string ct = "text/plain"; string postData = string.Format("?text={0}&from={1}&to={2}&contentType={3}", HttpUtility.UrlEncode(text), fromCulture, toCulture, HttpUtility.UrlEncode(ct)); web.Encoding = Encoding.UTF8; res = web.DownloadString(serviceUrl + postData); } catch (Exception e) { ErrorMessage = e.GetBaseException().Message; return null; } // result is a single XML Element fragment var doc = new XmlDocument(); doc.LoadXml(res); return doc.DocumentElement.InnerText; } The first of this code deals with ensuring the auth token exists. You can either pass the token into the method manually or let the method automatically retrieve the auth code on its own. In my case I'm using this inside of a Web application and in that situation I simply need to re-authenticate every time as there's no convenient way to manage the lifetime of the auth cookie. The auth token is added as an Authorization HTTP header prefixed with 'Bearer ' and attached to the request. The text to translate, the from and to language codes and a result format are passed on the query string of this HTTP GET request against the Translate API. The translate API returns an XML string which contains a single element with the translated string. Using the Wrapper Methods It should be pretty obvious how to use these two methods but here are a couple of test methods that demonstrate the two usage scenarios:[TestMethod] public void TranslateBingWithAuthTest() { var translate = new TranslationServices(); string clientId = DbResourceConfiguration.Current.BingClientId; string clientSecret = DbResourceConfiguration.Current.BingClientSecret; string auth = translate.GetBingAuthToken(clientId, clientSecret); Assert.IsNotNull(auth); string text = translate.TranslateBing("Hello World we're back home!", "en", "de",auth); Assert.IsNotNull(text, translate.ErrorMessage); Console.WriteLine(text); } [TestMethod] public void TranslateBingIntegratedTest() { var translate = new TranslationServices(); string text = translate.TranslateBing("Hello World we're back home!","en","de"); Assert.IsNotNull(text, translate.ErrorMessage); Console.WriteLine(text); } Other API Methods The Translate API has a number of methods available and this one is the simplest one but probably also the most common one that translates a single string. You can find additional methods for this API here: http://msdn.microsoft.com/en-us/library/ff512419.aspx Soap and AJAX APIs are also available and documented on MSDN: http://msdn.microsoft.com/en-us/library/dd576287.aspx These links will be your starting points for calling other methods in this API. Dual Interface I've talked about my database driven localization provider here in the past, and it's for this tool that I added the Bing localization support. Basically I have a localization administration form that allows me to translate individual strings right out of the UI, using both Google and Bing APIs: As you can see in this example, the results from Google and Bing can vary quite a bit - in this case Google is stumped while Bing actually generated a valid translation. At other times it's the other way around - it's pretty useful to see multiple translations at the same time. Here I can choose from one of the values and driectly embed them into the translated text field. Lost in Translation There you have it. As I mentioned using the API once you have all the bureaucratic crap out of the way calling the APIs is fairly straight forward and reasonably fast, even if you have to call the Auth API for every call. Hopefully this post will help out a few of you trying to navigate the Microsoft bureaucracy, at least until next time Microsoft upends everything and introduces new ways to sign up again. Until then - happy translating… Related Posts Translation method Source on Github Translating with Google Translate without Google API Keys Creating a data-driven ASP.NET Resource Provider© Rick Strahl, West Wind Technologies, 2005-2013Posted in Localization ASP.NET .NET Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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Getting Started Building Windows 8 Store Apps with XAML/C#

- by dwahlin

Technology is fun isn’t it? As soon as you think you’ve figured out where things are heading a new technology comes onto the scene, changes things up, and offers new opportunities. One of the new technologies I’ve been spending quite a bit of time with lately is Windows 8 store applications. I posted my thoughts about Windows 8 during the BUILD conference in 2011 and still feel excited about the opportunity there. Time will tell how well it ends up being accepted by consumers but I’m hopeful that it’ll take off. I currently have two Windows 8 store application concepts I’m working on with one being built in XAML/C# and another in HTML/JavaScript. I really like that Microsoft supports both options since it caters to a variety of developers and makes it easy to get started regardless if you’re a desktop developer or Web developer. Here’s a quick look at how the technologies are organized in Windows 8: In this post I’ll focus on the basics of Windows 8 store XAML/C# apps by looking at features, files, and code provided by Visual Studio projects. To get started building these types of apps you’ll definitely need to have some knowledge of XAML and C#. Let’s get started by looking at the Windows 8 store project types available in Visual Studio 2012. Windows 8 Store XAML/C# Project Types When you open Visual Studio 2012 you’ll see a new entry under C# named Windows Store. It includes 6 different project types as shown next. The Blank App project provides initial starter code and a single page whereas the Grid App and Split App templates provide quite a bit more code as well as multiple pages for your application. The other projects available can be be used to create a class library project that runs in Windows 8 store apps, a WinRT component such as a custom control, and a unit test library project respectively. If you’re building an application that displays data in groups using the “tile” concept then the Grid App or Split App project templates are a good place to start. An example of the initial screens generated by each project is shown next: Grid App Split View App When a user clicks a tile in a Grid App they can view details about the tile data. With a Split View app groups/categories are shown and when the user clicks on a group they can see a list of all the different items and then drill-down into them: For the remainder of this post I’ll focus on functionality provided by the Blank App project since it provides a simple way to get started learning the fundamentals of building Windows 8 store apps. Blank App Project Walkthrough The Blank App project is a great place to start since it’s simple and lets you focus on the basics. In this post I’ll focus on what it provides you out of the box and cover additional details in future posts. Once you have the basics down you can move to the other project types if you need the functionality they provide. The Blank App project template does exactly what it says – you get an empty project with a few starter files added to help get you going. This is a good option if you’ll be building an app that doesn’t fit into the grid layout view that you see a lot of Windows 8 store apps following (such as on the Windows 8 start screen). I ended up starting with the Blank App project template for the app I’m currently working on since I’m not displaying data/image tiles (something the Grid App project does well) or drilling down into lists of data (functionality that the Split App project provides). The Blank App project provides images for the tiles and splash screen (you’ll definitely want to change these), a StandardStyles.xaml resource dictionary that includes a lot of helpful styles such as buttons for the AppBar (a special type of menu in Windows 8 store apps), an App.xaml file, and the app’s main page which is named MainPage.xaml. It also adds a Package.appxmanifest that is used to define functionality that your app requires, app information used in the store, plus more. The App.xaml, App.xaml.cs and StandardStyles.xaml Files The App.xaml file handles loading a resource dictionary named StandardStyles.xaml which has several key styles used throughout the application: <Application x:Class="BlankApp.App" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:local="using:BlankApp"> <Application.Resources> <ResourceDictionary> <ResourceDictionary.MergedDictionaries>  <ResourceDictionary Source="Common/StandardStyles.xaml"/> </ResourceDictionary.MergedDictionaries> </ResourceDictionary> </Application.Resources> </Application> StandardStyles.xaml has style definitions for different text styles and AppBar buttons. If you scroll down toward the middle of the file you’ll see that many AppBar button styles are included such as one for an edit icon. Button styles like this can be used to quickly and easily add icons/buttons into your application without having to be an expert in design. <Style x:Key="EditAppBarButtonStyle" TargetType="ButtonBase" BasedOn="{StaticResource AppBarButtonStyle}"> <Setter Property="AutomationProperties.AutomationId" Value="EditAppBarButton"/> <Setter Property="AutomationProperties.Name" Value="Edit"/> <Setter Property="Content" Value=""/> </Style> Switching over to App.xaml.cs, it includes some code to help get you started. An OnLaunched() method is added to handle creating a Frame that child pages such as MainPage.xaml can be loaded into. The Frame has the same overall purpose as the one found in WPF and Silverlight applications - it’s used to navigate between pages in an application. /// <summary> /// Invoked when the application is launched normally by the end user. Other entry points /// will be used when the application is launched to open a specific file, to display /// search results, and so forth. /// </summary> /// <param name="args">Details about the launch request and process.</param> protected override void OnLaunched(LaunchActivatedEventArgs args) { Frame rootFrame = Window.Current.Content as Frame; // Do not repeat app initialization when the Window already has content, // just ensure that the window is active if (rootFrame == null) { // Create a Frame to act as the navigation context and navigate to the first page rootFrame = new Frame(); if (args.PreviousExecutionState == ApplicationExecutionState.Terminated) { //TODO: Load state from previously suspended application } // Place the frame in the current Window Window.Current.Content = rootFrame; } if (rootFrame.Content == null) { // When the navigation stack isn't restored navigate to the first page, // configuring the new page by passing required information as a navigation // parameter if (!rootFrame.Navigate(typeof(MainPage), args.Arguments)) { throw new Exception("Failed to create initial page"); } } // Ensure the current window is active Window.Current.Activate(); } Notice that in addition to creating a Frame the code also checks to see if the app was previously terminated so that you can load any state/data that the user may need when the app is launched again. If you’re new to the lifecycle of Windows 8 store apps the following image shows how an app can be running, suspended, and terminated. If the user switches from an app they’re running the app will be suspended in memory. The app may stay suspended or may be terminated depending on how much memory the OS thinks it needs so it’s important to save state in case the application is ultimately terminated and has to be started fresh. Although I won’t cover saving application state here, additional information can be found at http://msdn.microsoft.com/en-us/library/windows/apps/xaml/hh465099.aspx. Another method in App.xaml.cs named OnSuspending() is also included in App.xaml.cs that can be used to store state as the user switches to another application: /// <summary> /// Invoked when application execution is being suspended. Application state is saved /// without knowing whether the application will be terminated or resumed with the contents /// of memory still intact. /// </summary> /// <param name="sender">The source of the suspend request.</param> /// <param name="e">Details about the suspend request.</param> private void OnSuspending(object sender, SuspendingEventArgs e) { var deferral = e.SuspendingOperation.GetDeferral(); //TODO: Save application state and stop any background activity deferral.Complete(); } The MainPage.xaml and MainPage.xaml.cs Files The Blank App project adds a file named MainPage.xaml that acts as the initial screen for the application. It doesn’t include anything aside from an empty <Grid> XAML element in it. The code-behind class named MainPage.xaml.cs includes a constructor as well as a method named OnNavigatedTo() that is called once the page is displayed in the frame. /// <summary> /// An empty page that can be used on its own or navigated to within a Frame. /// </summary> public sealed partial class MainPage : Page { public MainPage() { this.InitializeComponent(); } /// <summary> /// Invoked when this page is about to be displayed in a Frame. /// </summary> /// <param name="e">Event data that describes how this page was reached. The Parameter /// property is typically used to configure the page.</param> protected override void OnNavigatedTo(NavigationEventArgs e) { } } If you’re experienced with XAML you can switch to Design mode and start dragging and dropping XAML controls from the ToolBox in Visual Studio. If you prefer to type XAML you can do that as well in the XAML editor or while in split mode. Many of the controls available in WPF and Silverlight are included such as Canvas, Grid, StackPanel, and Border for layout. Standard input controls are also included such as TextBox, CheckBox, PasswordBox, RadioButton, ComboBox, ListBox, and more. MediaElement is available for rendering video or playing audio files. Some of the “common” XAML controls included out of the box are shown next: Although XAML/C# Windows 8 store apps don’t include all of the functionality available in Silverlight 5, the core functionality required to build store apps is there with additional functionality available in open source projects such as Callisto (started by Microsoft’s Tim Heuer), Q42.WinRT, and others. Standard XAML data binding can be used to bind C# objects to controls, converters can be used to manipulate data during the data binding process, and custom styles and templates can be applied to controls to modify them. Although Visual Studio 2012 doesn’t support visually creating styles or templates, Expression Blend 5 handles that very well. To get started building the initial screen of a Windows 8 app you can start adding controls as mentioned earlier. Simply place them inside of the <Grid> element that’s included. You can arrange controls in a stacked manner using the StackPanel control, add a border around controls using the Border control, arrange controls in columns and rows using the Grid control, or absolutely position controls using the Canvas control. One of the controls that may be new to you is the AppBar. It can be used to add menu/toolbar functionality into a store app and keep the app clean and focused. You can place an AppBar at the top or bottom of the screen. A user on a touch device can swipe up to display the bottom AppBar or right-click when using a mouse. An example of defining an AppBar that contains an Edit button is shown next. The EditAppBarButtonStyle is available in the StandardStyles.xaml file mentioned earlier. <Page.BottomAppBar> <AppBar x:Name="ApplicationAppBar" Padding="10,0,10,0" AutomationProperties.Name="Bottom App Bar"> <Grid> <StackPanel x:Name="RightPanel" Orientation="Horizontal" Grid.Column="1" HorizontalAlignment="Right"> <Button x:Name="Edit" Style="{StaticResource EditAppBarButtonStyle}" Tag="Edit" /> </StackPanel> </Grid> </AppBar> </Page.BottomAppBar> Like standard XAML controls, the <Button> control in the AppBar can be wired to an event handler method in the MainPage.Xaml.cs file or even bound to a ViewModel object using “commanding” if your app follows the Model-View-ViewModel (MVVM) pattern (check out the MVVM Light package available through NuGet if you’re using MVVM with Windows 8 store apps). The AppBar can be used to navigate to different screens, show and hide controls, display dialogs, show settings screens, and more. The Package.appxmanifest File The Package.appxmanifest file contains configuration details about your Windows 8 store app. By double-clicking it in Visual Studio you can define the splash screen image, small and wide logo images used for tiles on the start screen, orientation information, and more. You can also define what capabilities the app has such as if it uses the Internet, supports geolocation functionality, requires a microphone or webcam, etc. App declarations such as background processes, file picker functionality, and sharing can also be defined Finally, information about how the app is packaged for deployment to the store can also be defined. Summary If you already have some experience working with XAML technologies you’ll find that getting started building Windows 8 applications is pretty straightforward. Many of the controls available in Silverlight and WPF are available making it easy to get started without having to relearn a lot of new technologies. In the next post in this series I’ll discuss additional features that can be used in your Windows 8 store apps.

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OpenVPN not connecting

- by LandArch

There have been a number of post similar to this, but none seem to satisfy my need. Plus I am a Ubuntu newbie. I followed this tutorial to completely set up OpenVPN on Ubuntu 12.04 server. Here is my server.conf file ################################################# # Sample OpenVPN 2.0 config file for # # multi-client server. # # # # This file is for the server side # # of a many-clients <-> one-server # # OpenVPN configuration. # # # # OpenVPN also supports # # single-machine <-> single-machine # # configurations (See the Examples page # # on the web site for more info). # # # # This config should work on Windows # # or Linux/BSD systems. Remember on # # Windows to quote pathnames and use # # double backslashes, e.g.: # # "C:\\Program Files\\OpenVPN\\config\\foo.key" # # # # Comments are preceded with '#' or ';' # ################################################# # Which local IP address should OpenVPN # listen on? (optional) local 192.168.13.8 # Which TCP/UDP port should OpenVPN listen on? # If you want to run multiple OpenVPN instances # on the same machine, use a different port # number for each one. You will need to # open up this port on your firewall. port 1194 # TCP or UDP server? proto tcp ;proto udp # "dev tun" will create a routed IP tunnel, # "dev tap" will create an ethernet tunnel. # Use "dev tap0" if you are ethernet bridging # and have precreated a tap0 virtual interface # and bridged it with your ethernet interface. # If you want to control access policies # over the VPN, you must create firewall # rules for the the TUN/TAP interface. # On non-Windows systems, you can give # an explicit unit number, such as tun0. # On Windows, use "dev-node" for this. # On most systems, the VPN will not function # unless you partially or fully disable # the firewall for the TUN/TAP interface. dev tap0 up "/etc/openvpn/up.sh br0" down "/etc/openvpn/down.sh br0" ;dev tun # Windows needs the TAP-Win32 adapter name # from the Network Connections panel if you # have more than one. On XP SP2 or higher, # you may need to selectively disable the # Windows firewall for the TAP adapter. # Non-Windows systems usually don't need this. ;dev-node MyTap # SSL/TLS root certificate (ca), certificate # (cert), and private key (key). Each client # and the server must have their own cert and # key file. The server and all clients will # use the same ca file. # # See the "easy-rsa" directory for a series # of scripts for generating RSA certificates # and private keys. Remember to use # a unique Common Name for the server # and each of the client certificates. # # Any X509 key management system can be used. # OpenVPN can also use a PKCS #12 formatted key file # (see "pkcs12" directive in man page). ca "/etc/openvpn/ca.crt" cert "/etc/openvpn/server.crt" key "/etc/openvpn/server.key" # This file should be kept secret # Diffie hellman parameters. # Generate your own with: # openssl dhparam -out dh1024.pem 1024 # Substitute 2048 for 1024 if you are using # 2048 bit keys. dh dh1024.pem # Configure server mode and supply a VPN subnet # for OpenVPN to draw client addresses from. # The server will take 10.8.0.1 for itself, # the rest will be made available to clients. # Each client will be able to reach the server # on 10.8.0.1. Comment this line out if you are # ethernet bridging. See the man page for more info. ;server 10.8.0.0 255.255.255.0 # Maintain a record of client <-> virtual IP address # associations in this file. If OpenVPN goes down or # is restarted, reconnecting clients can be assigned # the same virtual IP address from the pool that was # previously assigned. ifconfig-pool-persist ipp.txt # Configure server mode for ethernet bridging. # You must first use your OS's bridging capability # to bridge the TAP interface with the ethernet # NIC interface. Then you must manually set the # IP/netmask on the bridge interface, here we # assume 10.8.0.4/255.255.255.0. Finally we # must set aside an IP range in this subnet # (start=10.8.0.50 end=10.8.0.100) to allocate # to connecting clients. Leave this line commented # out unless you are ethernet bridging. server-bridge 192.168.13.101 255.255.255.0 192.168.13.105 192.168.13.200 # Configure server mode for ethernet bridging # using a DHCP-proxy, where clients talk # to the OpenVPN server-side DHCP server # to receive their IP address allocation # and DNS server addresses. You must first use # your OS's bridging capability to bridge the TAP # interface with the ethernet NIC interface. # Note: this mode only works on clients (such as # Windows), where the client-side TAP adapter is # bound to a DHCP client. ;server-bridge # Push routes to the client to allow it # to reach other private subnets behind # the server. Remember that these # private subnets will also need # to know to route the OpenVPN client # address pool (10.8.0.0/255.255.255.0) # back to the OpenVPN server. push "route 192.168.13.1 255.255.255.0" push "dhcp-option DNS 192.168.13.201" push "dhcp-option DOMAIN blahblah.dyndns-wiki.com" ;push "route 192.168.20.0 255.255.255.0" # To assign specific IP addresses to specific # clients or if a connecting client has a private # subnet behind it that should also have VPN access, # use the subdirectory "ccd" for client-specific # configuration files (see man page for more info). # EXAMPLE: Suppose the client # having the certificate common name "Thelonious" # also has a small subnet behind his connecting # machine, such as 192.168.40.128/255.255.255.248. # First, uncomment out these lines: ;client-config-dir ccd ;route 192.168.40.128 255.255.255.248 # Then create a file ccd/Thelonious with this line: # iroute 192.168.40.128 255.255.255.248 # This will allow Thelonious' private subnet to # access the VPN. This example will only work # if you are routing, not bridging, i.e. you are # using "dev tun" and "server" directives. # EXAMPLE: Suppose you want to give # Thelonious a fixed VPN IP address of 10.9.0.1. # First uncomment out these lines: ;client-config-dir ccd ;route 10.9.0.0 255.255.255.252 # Then add this line to ccd/Thelonious: # ifconfig-push 10.9.0.1 10.9.0.2 # Suppose that you want to enable different # firewall access policies for different groups # of clients. There are two methods: # (1) Run multiple OpenVPN daemons, one for each # group, and firewall the TUN/TAP interface # for each group/daemon appropriately. # (2) (Advanced) Create a script to dynamically # modify the firewall in response to access # from different clients. See man # page for more info on learn-address script. ;learn-address ./script # If enabled, this directive will configure # all clients to redirect their default # network gateway through the VPN, causing # all IP traffic such as web browsing and # and DNS lookups to go through the VPN # (The OpenVPN server machine may need to NAT # or bridge the TUN/TAP interface to the internet # in order for this to work properly). ;push "redirect-gateway def1 bypass-dhcp" # Certain Windows-specific network settings # can be pushed to clients, such as DNS # or WINS server addresses. CAVEAT: # http://openvpn.net/faq.html#dhcpcaveats # The addresses below refer to the public # DNS servers provided by opendns.com. ;push "dhcp-option DNS 208.67.222.222" ;push "dhcp-option DNS 208.67.220.220" # Uncomment this directive to allow different # clients to be able to "see" each other. # By default, clients will only see the server. # To force clients to only see the server, you # will also need to appropriately firewall the # server's TUN/TAP interface. ;client-to-client # Uncomment this directive if multiple clients # might connect with the same certificate/key # files or common names. This is recommended # only for testing purposes. For production use, # each client should have its own certificate/key # pair. # # IF YOU HAVE NOT GENERATED INDIVIDUAL # CERTIFICATE/KEY PAIRS FOR EACH CLIENT, # EACH HAVING ITS OWN UNIQUE "COMMON NAME", # UNCOMMENT THIS LINE OUT. ;duplicate-cn # The keepalive directive causes ping-like # messages to be sent back and forth over # the link so that each side knows when # the other side has gone down. # Ping every 10 seconds, assume that remote # peer is down if no ping received during # a 120 second time period. keepalive 10 120 # For extra security beyond that provided # by SSL/TLS, create an "HMAC firewall" # to help block DoS attacks and UDP port flooding. # # Generate with: # openvpn --genkey --secret ta.key # # The server and each client must have # a copy of this key. # The second parameter should be '0' # on the server and '1' on the clients. ;tls-auth ta.key 0 # This file is secret # Select a cryptographic cipher. # This config item must be copied to # the client config file as well. ;cipher BF-CBC # Blowfish (default) ;cipher AES-128-CBC # AES ;cipher DES-EDE3-CBC # Triple-DES # Enable compression on the VPN link. # If you enable it here, you must also # enable it in the client config file. comp-lzo # The maximum number of concurrently connected # clients we want to allow. ;max-clients 100 # It's a good idea to reduce the OpenVPN # daemon's privileges after initialization. # # You can uncomment this out on # non-Windows systems. user nobody group nogroup # The persist options will try to avoid # accessing certain resources on restart # that may no longer be accessible because # of the privilege downgrade. persist-key persist-tun # Output a short status file showing # current connections, truncated # and rewritten every minute. status openvpn-status.log # By default, log messages will go to the syslog (or # on Windows, if running as a service, they will go to # the "\Program Files\OpenVPN\log" directory). # Use log or log-append to override this default. # "log" will truncate the log file on OpenVPN startup, # while "log-append" will append to it. Use one # or the other (but not both). ;log openvpn.log ;log-append openvpn.log # Set the appropriate level of log # file verbosity. # # 0 is silent, except for fatal errors # 4 is reasonable for general usage # 5 and 6 can help to debug connection problems # 9 is extremely verbose verb 3 # Silence repeating messages. At most 20 # sequential messages of the same message # category will be output to the log. ;mute 20 I am using Windows 7 as the Client and set that up accordingly using the OpenVPN GUI. That conf file is as follows: ############################################## # Sample client-side OpenVPN 2.0 config file # # for connecting to multi-client server. # # # # This configuration can be used by multiple # # clients, however each client should have # # its own cert and key files. # # # # On Windows, you might want to rename this # # file so it has a .ovpn extension # ############################################## # Specify that we are a client and that we # will be pulling certain config file directives # from the server. client # Use the same setting as you are using on # the server. # On most systems, the VPN will not function # unless you partially or fully disable # the firewall for the TUN/TAP interface. dev tap0 up "/etc/openvpn/up.sh br0" down "/etc/openvpn/down.sh br0" ;dev tun # Windows needs the TAP-Win32 adapter name # from the Network Connections panel # if you have more than one. On XP SP2, # you may need to disable the firewall # for the TAP adapter. ;dev-node MyTap # Are we connecting to a TCP or # UDP server? Use the same setting as # on the server. proto tcp ;proto udp # The hostname/IP and port of the server. # You can have multiple remote entries # to load balance between the servers. blahblah.dyndns-wiki.com 1194 ;remote my-server-2 1194 # Choose a random host from the remote # list for load-balancing. Otherwise # try hosts in the order specified. ;remote-random # Keep trying indefinitely to resolve the # host name of the OpenVPN server. Very useful # on machines which are not permanently connected # to the internet such as laptops. resolv-retry infinite # Most clients don't need to bind to # a specific local port number. nobind # Downgrade privileges after initialization (non-Windows only) user nobody group nobody # Try to preserve some state across restarts. persist-key persist-tun # If you are connecting through an # HTTP proxy to reach the actual OpenVPN # server, put the proxy server/IP and # port number here. See the man page # if your proxy server requires # authentication. ;http-proxy-retry # retry on connection failures ;http-proxy [proxy server] [proxy port #] # Wireless networks often produce a lot # of duplicate packets. Set this flag # to silence duplicate packet warnings. ;mute-replay-warnings # SSL/TLS parms. # See the server config file for more # description. It's best to use # a separate .crt/.key file pair # for each client. A single ca # file can be used for all clients. ca "C:\\Program Files\OpenVPN\config\\ca.crt" cert "C:\\Program Files\OpenVPN\config\\ChadMWade-THINK.crt" key "C:\\Program Files\OpenVPN\config\\ChadMWade-THINK.key" # Verify server certificate by checking # that the certicate has the nsCertType # field set to "server". This is an # important precaution to protect against # a potential attack discussed here: # http://openvpn.net/howto.html#mitm # # To use this feature, you will need to generate # your server certificates with the nsCertType # field set to "server". The build-key-server # script in the easy-rsa folder will do this. ns-cert-type server # If a tls-auth key is used on the server # then every client must also have the key. ;tls-auth ta.key 1 # Select a cryptographic cipher. # If the cipher option is used on the server # then you must also specify it here. ;cipher x # Enable compression on the VPN link. # Don't enable this unless it is also # enabled in the server config file. comp-lzo # Set log file verbosity. verb 3 # Silence repeating messages ;mute 20 Not sure whats left to do.

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Mobile Friendly Websites with CSS Media Queries

- by dwahlin

In a previous post the concept of CSS media queries was introduced and I discussed the fundamentals of how they can be used to target different screen sizes. I showed how they could be used to convert a 3-column wide page into a more vertical view of data that displays better on devices such as an iPhone: In this post I'll provide an additional look at how CSS media queries can be used to mobile-enable a sample site called "Widget Masters" without having to change any server-side code or HTML code. The site that will be discussed is shown next: This site has some of the standard items shown in most websites today including a title area, menu bar, and sections where data is displayed. Without including CSS media queries the site is readable but has to be zoomed out to see everything on a mobile device, cuts-off some of the menu items, and requires horizontal scrolling to get to additional content. The following image shows what the site looks like on an iPhone. While the site works on mobile devices it's definitely not optimized for mobile. Let's take a look at how CSS media queries can be used to override existing styles in the site based on different screen widths. Adding CSS Media Queries into a Site The Widget Masters Website relies on standard CSS combined with HTML5 elements to provide the layout shown earlier. For example, to layout the menu bar shown at the top of the page the nav element is used as shown next. A standard div element could certainly be used as well if desired. <nav> <ul class="clearfix"> <li><a href="#home">Home</a></li> <li><a href="#products">Products</a></li> <li><a href="#aboutus">About Us</a></li> <li><a href="#contactus">Contact Us</a></li> <li><a href="#store">Store</a></li> </ul> </nav> This HTML is combined with the CSS shown next to add a CSS3 gradient, handle the horizontal orientation, and add some general hover effects. nav { width: 100%; } nav ul { border-radius: 6px; height: 40px; width: 100%; margin: 0; padding: 0; background: rgb(125,126,125); /* Old browsers */ background: -moz-linear-gradient(top, rgba(125,126,125,1) 0%, rgba(14,14,14,1) 100%); /* FF3.6+ */ background: -webkit-gradient(linear, left top, left bottom, color-stop(0%,rgba(125,126,125,1)), color-stop(100%,rgba(14,14,14,1))); /* Chrome,Safari4+ */ background: -webkit-linear-gradient(top, rgba(125,126,125,1) 0%, rgba(14,14,14,1) 100%); /* Chrome10+,Safari5.1+ */ background: -o-linear-gradient(top, rgba(125,126,125,1) 0%, rgba(14,14,14,1) 100%); /* Opera 11.10+ */ background: -ms-linear-gradient(top, rgba(125,126,125,1) 0%, rgba(14,14,14,1) 100%); /* IE10+ */ background: linear-gradient(top, rgba(125,126,125,1) 0%, rgba(14,14,14,1) 100%); /* W3C */ filter: progid:DXImageTransform.Microsoft.gradient( startColorstr='#7d7e7d', endColorstr='#0e0e0e',GradientType=0 ); /* IE6-9 */ } nav ul > li { list-style: none; float: left; margin: 0; padding: 0; } nav ul > li:first-child { margin-left: 8px; } nav ul > li > a { color: #ccc; text-decoration: none; line-height: 2.8em; font-size: 0.95em; font-weight: bold; padding: 8px 25px 7px 25px; font-family: Arial, Helvetica, sans-serif; } nav ul > li a:hover { background-color: rgba(0, 0, 0, 0.1); color: #fff; } When mobile devices hit the site the layout of the menu items needs to be adjusted so that they're all visible without having to swipe left or right to get to them. This type of modification can be accomplished using CSS media queries by targeting specific screen sizes. To start, a media query can be added into the site's CSS file as shown next: @media screen and (max-width:320px) { /* CSS style overrides for this screen width go here */ } This media query targets screens that have a maximum width of 320 pixels. Additional types of queries can also be added – refer to my previous post for more details as well as resources that can be used to test media queries in different devices. In that post I emphasize (and I'll emphasize again) that CSS media queries only modify the overall layout and look and feel of a site. They don't optimize the site as far as the size of the images or content sent to the device which is important to keep in mind. To make the navigation menu more accessible on devices such as an iPhone or Android the CSS shown next can be used. This code changes the height of the menu from 40 pixels to 100%, takes off the li element floats, changes the line-height, and changes the margins. @media screen and (max-width:320px) { nav ul { height: 100%; } nav ul > li { float: none; } nav ul > li a { line-height: 1.5em; } nav ul > li:first-child { margin-left: 0px; } /* Additional CSS overrides go here */ } The following image shows an example of what the menu look like when run on a device with a width of 320 pixels: Mobile devices with a maximum width of 480 pixels need different CSS styles applied since they have 160 additional pixels of width. This can be done by adding a new CSS media query into the stylesheet as shown next. Looking through the CSS you'll see that only a minimal override is added to adjust the padding of anchor tags since the menu fits by default in this screen width. @media screen and (max-width: 480px) { nav ul > li > a { padding: 8px 10px 7px 10px; } } Running the site on a device with 480 pixels results in the menu shown next being rendered. Notice that the space between the menu items is much smaller compared to what was shown when the main site loads in a standard browser. In addition to modifying the menu, the 3 horizontal content sections shown earlier can be changed from a horizontal layout to a vertical layout so that they look good on a variety of smaller mobile devices and are easier to navigate by end users. The HTML5 article and section elements are used as containers for the 3 sections in the site as shown next: <article class="clearfix"> <section id="info"> <header>Why Choose Us?</header> <br /> <img id="mainImage" src="Images/ArticleImage.png" title="Article Image" /> <p> Post emensos insuperabilis expeditionis eventus languentibus partium animis, quas periculorum varietas fregerat et laborum, nondum tubarum cessante clangore vel milite locato per stationes hibernas. </p> </section> <section id="products"> <header>Products</header> <br /> <img id="gearsImage" src="Images/Gears.png" title="Article Image" /> <p> <ul> <li>Widget 1</li> <li>Widget 2</li> <li>Widget 3</li> <li>Widget 4</li> <li>Widget 5</li> </ul> </p> </section> <section id="FAQ"> <header>FAQ</header> <br /> <img id="faqImage" src="Images/faq.png" title="Article Image" /> <p> <ul> <li>FAQ 1</li> <li>FAQ 2</li> <li>FAQ 3</li> <li>FAQ 4</li> <li>FAQ 5</li> </ul> </p> </section> </article> To force the sections into a vertical layout for smaller mobile devices the CSS styles shown next can be added into the media queries targeting 320 pixel and 480 pixel widths. Styles to target the display size of the images in each section are also included. It's important to note that the original image is still being downloaded from the server and isn't being optimized in any way for the mobile device. It's certainly possible for the CSS to include URL information for a mobile-optimized image if desired. @media screen and (max-width:320px) { section { float: none; width: 97%; margin: 0px; padding: 5px; } #wrapper { padding: 5px; width: 96%; } #mainImage, #gearsImage, #faqImage { width: 100%; height: 100px; } } @media screen and (max-width: 480px) { section { float: none; width: 98%; margin: 0px 0px 10px 0px; padding: 5px; } article > section:last-child { margin-right: 0px; float: none; } #bottomSection { width: 99%; } #wrapper { padding: 5px; width: 96%; } #mainImage, #gearsImage, #faqImage { width: 100%; height: 100px; } } The following images show the site rendered on an iPhone with the CSS media queries in place. Each of the sections now displays vertically making it much easier for the user to access them. Images inside of each section also scale appropriately to fit properly. CSS media queries provide a great way to override default styles in a website and target devices with different resolutions. In this post you've seen how CSS media queries can be used to convert a standard browser-based site into a site that is more accessible to mobile users. Although much more can be done to optimize sites for mobile, CSS media queries provide a nice starting point if you don't have the time or resources to create mobile-specific versions of sites.

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Visual Studio 2010 and .NET 4 Released

- by ScottGu

The final release of Visual Studio 2010 and .NET 4 is now available. Download and Install Today MSDN subscribers, as well as WebsiteSpark/BizSpark/DreamSpark members, can now download the final releases of Visual Studio 2010 and TFS 2010 through the MSDN subscribers download center. If you are not an MSDN Subscriber, you can download free 90-day trial editions of Visual Studio 2010. Or you can can download the free Visual Studio express editions of Visual Web Developer 2010, Visual Basic 2010, Visual C# 2010 and Visual C++. These express editions are available completely for free (and never time out). If you are looking for an easy way to setup a new machine for web-development you can automate installing ASP.NET 4, ASP.NET MVC 2, IIS, SQL Server Express and Visual Web Developer 2010 Express really quickly with the Microsoft Web Platform Installer (just click the install button on the page). What is new with VS 2010 and .NET 4 Today’s release is a big one – and brings with it a ton of new feature and capabilities. One of the things we tried hard to focus on with this release was to invest heavily in making existing applications, projects and developer experiences better. What this means is that you don’t need to read 1000+ page books or spend time learning major new concepts in order to take advantage of the release. There are literally thousands of improvements (both big and small) that make you more productive and successful without having to learn big new concepts in order to start using them. Below is just a small sampling of some of the improvements with this release: Visual Studio 2010 IDE Visual Studio 2010 now supports multiple-monitors (enabling much better use of screen real-estate). It has new code Intellisense support that makes it easier to find and use classes and methods. It has improved code navigation support for searching code-bases and seeing how code is called and used. It has new code visualization support that allows you to see the relationships across projects and classes within projects, as well as to automatically generate sequence diagrams to chart execution flow. The editor now supports HTML and JavaScript snippet support as well as improved JavaScript intellisense. The VS 2010 Debugger and Profiling support is now much, much richer and enables new features like Intellitrace (aka Historical Debugging), debugging of Crash/Dump files, and better parallel debugging. VS 2010’s multi-targeting support is now much richer, and enables you to use VS 2010 to target .NET 2, .NET 3, .NET 3.5 and .NET 4 applications. And the infamous Add Reference dialog now loads much faster. TFS 2010 is now easy to setup (you can now install the server in under 10 minutes) and enables great source-control, bug/work-item tracking, and continuous integration support. Testing (both automated and manual) is now much, much richer. And VS 2010 Premium and Ultimate provide much richer architecture and design tooling support. VB and C# Language Features VB and C# in VS 2010 both contain a bunch of new features and capabilities. VB adds new support for automatic properties, collection initializers, and implicit line continuation support among many other features. C# adds support for optional parameters and named arguments, a new dynamic keyword, co-variance and contra-variance, and among many other features. ASP.NET 4 and ASP.NET MVC 2 With ASP.NET 4, Web Forms controls now render clean, semantically correct, and CSS friendly HTML markup. Built-in URL routing functionality allows you to expose clean, search engine friendly, URLs and increase the traffic to your Website. ViewState within applications can now be more easily controlled and made smaller. ASP.NET Dynamic Data support has been expanded. More controls, including rich charting and data controls, are now built-into ASP.NET 4 and enable you to build applications even faster. New starter project templates now make it easier to get going with new projects. SEO enhancements make it easier to drive traffic to your public facing sites. And web.config files are now clean and simple. ASP.NET MVC 2 is now built-into VS 2010 and ASP.NET 4, and provides a great way to build web sites and applications using a model-view-controller based pattern. ASP.NET MVC 2 adds features to easily enable client and server validation logic, provides new strongly-typed HTML and UI-scaffolding helper methods. It also enables more modular/reusable applications. The new <%: %> syntax in ASP.NET makes it easier to HTML encode output. Visual Studio 2010 also now includes better tooling support for unit testing and TDD. In particular, “Consume first intellisense” and “generate from usage" support within VS 2010 make it easier to write your unit tests first, and then drive your implementation from them. Deploying ASP.NET applications gets a lot easier with this release. You can now publish your Websites and applications to a staging or production server from within Visual Studio itself. Visual Studio 2010 makes it easy to transfer all your files, code, configuration, database schema and data in one complete package. VS 2010 also makes it easy to manage separate web.config configuration files settings depending upon whether you are in debug, release, staging or production modes. WPF 4 and Silverlight 4 WPF 4 includes a ton of new improvements and capabilities including more built-in controls, richer graphics features (cached composition, pixel shader 3 support, layoutrounding, and animation easing functions), a much improved text stack (with crisper text rendering, custom dictionary support, and selection and caret brush options). WPF 4 also includes a bunch of support to enable you to take advantage of new Windows 7 features – including multi-touch and Windows 7 shell integration. Silverlight 4 will launch this week as well. You can watch my Silverlight 4 launch keynote streamed live Tuesday (April 13th) at 8am Pacific Time. Silverlight 4 includes a ton of new capabilities – including a bunch for making it possible to build great business applications and out of the browser applications. I’ll be doing a separate blog post later this week (once it is live on the web) that talks more about its capabilities. Visual Studio 2010 now includes great tooling support for both WPF and Silverlight. The new VS 2010 WPF and Silverlight designer makes it much easier to build client applications as well as build great line of business solutions, as well as integrate and bind with data. Tooling support for Silverlight 4 with the final release of Visual Studio 2010 will be available when Silverlight 4 releases to the web this week. SharePoint and Azure Visual Studio 2010 now includes built-in support for building SharePoint applications. You can now create, edit, build, and debug SharePoint applications directly within Visual Studio 2010. You can also now use SharePoint with TFS 2010. Support for creating Azure-hosted applications is also now included with VS 2010 – allowing you to build ASP.NET and WCF based applications and host them within the cloud. Data Access Data access has a lot of improvements coming to it with .NET 4. Entity Framework 4 includes a ton of new features and capabilities – including support for model first and POCO development, default support for lazy loading, built-in support for pluralization/singularization of table/property names within the VS 2010 designer, full support for all the LINQ operators, the ability to optionally expose foreign keys on model objects (useful for some stateless web scenarios), disconnected API support to better handle N-Tier and stateless web scenarios, and T4 template customization support within VS 2010 to allow you to customize and automate how code is generated for you by the data designer. In addition to improvements with the Entity Framework, LINQ to SQL with .NET 4 also includes a bunch of nice improvements. WCF and Workflow WCF includes a bunch of great new capabilities – including better REST, activation and configuration support. WCF Data Services (formerly known as Astoria) and WCF RIA Services also now enable you to easily expose and work with data from remote clients. Windows Workflow is now much faster, includes flowchart services, and now makes it easier to make custom services than before. More details can be found here. CLR and Core .NET Library Improvements .NET 4 includes the new CLR 4 engine – which includes a lot of nice performance and feature improvements. CLR 4 engine now runs side-by-side in-process with older versions of the CLR – allowing you to use two different versions of .NET within the same process. It also includes improved COM interop support. The .NET 4 base class libraries (BCL) include a bunch of nice additions and refinements. In particular, the .NET 4 BCL now includes new parallel programming support that makes it much easier to build applications that take advantage of multiple CPUs and cores on a computer. This work dove-tails nicely with the new VS 2010 parallel debugger (making it much easier to debug parallel applications), as well as the new F# functional language support now included in the VS 2010 IDE. .NET 4 also now also has the Dynamic Language Runtime (DLR) library built-in – which makes it easier to use dynamic language functionality with .NET. MEF – a really cool library that enables rich extensibility – is also now built-into .NET 4 and included as part of the base class libraries. .NET 4 Client Profile The download size of the .NET 4 redist is now much smaller than it was before (the x86 full .NET 4 package is about 36MB). We also now have a .NET 4 Client Profile package which is a pure sub-set of the full .NET that can be used to streamline client application installs. C++ VS 2010 includes a bunch of great improvements for C++ development. This includes better C++ Intellisense support, MSBuild support for projects, improved parallel debugging and profiler support, MFC improvements, and a number of language features and compiler optimizations. My VS 2010 and .NET 4 Blog Series I’ve been cranking away on a blog series the last few months that highlights many of the new VS 2010 and .NET 4 improvements. The good news is that I have about 20 in-depth posts already written. The bad news (for me) is that I have about 200 more to go until I’m done! I’m going to try and keep adding a few more each week over the next few months to discuss the new improvements and how best to take advantage of them. Below is a list of the already written ones that you can check out today: Clean Web.Config Files Starter Project Templates Multi-targeting Multiple Monitor Support New Code Focused Web Profile Option HTML / ASP.NET / JavaScript Code Snippets Auto-Start ASP.NET Applications URL Routing with ASP.NET 4 Web Forms Searching and Navigating Code in VS 2010 VS 2010 Code Intellisense Improvements WPF 4 Add Reference Dialog Improvements SEO Improvements with ASP.NET 4 Output Cache Extensibility with ASP.NET 4 Built-in Charting Controls for ASP.NET and Windows Forms Cleaner HTML Markup with ASP.NET 4 - Client IDs Optional Parameters and Named Arguments in C# 4 - and a cool scenarios with ASP.NET MVC 2 Automatic Properties, Collection Initializers and Implicit Line Continuation Support with VB 2010 New <%: %> Syntax for HTML Encoding Output using ASP.NET 4 JavaScript Intellisense Improvements with VS 2010 Stay tuned to my blog as I post more. Also check out this page which links to a bunch of great articles and videos done by others. VS 2010 Installation Notes If you have installed a previous version of VS 2010 on your machine (either the beta or the RC) you must first uninstall it before installing the final VS 2010 release. I also recommend uninstalling .NET 4 betas (including both the client and full .NET 4 installs) as well as the other installs that come with VS 2010 (e.g. ASP.NET MVC 2 preview builds, etc). The uninstalls of the betas/RCs will clean up all the old state on your machine – after which you can install the final VS 2010 version and should have everything just work (this is what I’ve done on all of my machines and I haven’t had any problems). The VS 2010 and .NET 4 installs add a bunch of new managed assemblies to your machine. Some of these will be “NGEN’d” to native code during the actual install process (making them run fast). To avoid adding too much time to VS setup, though, we don’t NGEN all assemblies immediately – and instead will NGEN the rest in the background when your machine is idle. Until it finishes NGENing the assemblies they will be JIT’d to native code the first time they are used in a process – which for large assemblies can sometimes cause a slight performance hit. If you run into this you can manually force all assemblies to be NGEN’d to native code immediately (and not just wait till the machine is idle) by launching the Visual Studio command line prompt from the Windows Start Menu (Microsoft Visual Studio 2010->Visual Studio Tools->Visual Studio Command Prompt). Within the command prompt type “Ngen executequeueditems” – this will cause everything to be NGEN’d immediately. How to Buy Visual Studio 2010 You can can download and use the free Visual Studio express editions of Visual Web Developer 2010, Visual Basic 2010, Visual C# 2010 and Visual C++. These express editions are available completely for free (and never time out). You can buy a new copy of VS 2010 Professional that includes a 1 year subscription to MSDN Essentials for $799. MSDN Essentials includes a developer license of Windows 7 Ultimate, Windows Server 2008 R2 Enterprise, SQL Server 2008 DataCenter R2, and 20 hours of Azure hosting time. Subscribers also have access to MSDN’s Online Concierge, and Priority Support in MSDN Forums. Upgrade prices from previous releases of Visual Studio are also available. Existing Visual Studio 2005/2008 Standard customers can upgrade to Visual Studio 2010 Professional for a special $299 retail price until October. You can take advantage of this VS Standard->Professional upgrade promotion here. Web developers who build applications for others, and who are either independent developers or who work for companies with less than 10 employees, can also optionally take advantage of the Microsoft WebSiteSpark program. This program gives you three copies of Visual Studio 2010 Professional, 1 copy of Expression Studio, and 4 CPU licenses of both Windows 2008 R2 Web Server and SQL 2008 Web Edition that you can use to both develop and deploy applications with at no cost for 3 years. At the end of the 3 years there is no obligation to buy anything. You can sign-up for WebSiteSpark today in under 5 minutes – and immediately have access to the products to download. Summary Today’s release is a big one – and has a bunch of improvements for pretty much every developer. Thank you everyone who provided feedback, suggestions and reported bugs throughout the development process – we couldn’t have delivered it without you. Hope this helps, Scott P.S. In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu

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URL Rewrite – Multiple domains under one site. Part II

- by OWScott

I believe I have it … I’ve been meaning to put together the ultimate outgoing rule for hosting multiple domains under one site. I finally sat down this week and setup a few test cases, and created one rule to rule them all. In Part I of this two part series, I covered the incoming rule necessary to host a site in a subfolder of a website, while making it appear as if it’s in the root of the site. Part II won’t work without applying Part I first, so if you haven’t read it, I encourage you to read it now. However, the incoming rule by itself doesn’t address everything. Here’s the problem … Let’s say that we host www.site2.com in a subfolder called site2, off of masterdomain.com. This is the same example I used in Part I. Using an incoming rewrite rule, we are able to make a request to www.site2.com even though the site is really in the /site2 folder. The gotcha comes with any type of path that ASP.NET generates (I’m sure other scripting technologies could do the same too). ASP.NET thinks that the path to the root of the site is /site2, but the URL is /. See the issue? If ASP.NET generates a path or a redirect for us, it will always add /site2 to the URL. That results in a path that looks something like www.site2.com/site2. In Part I, I mentioned that you should add a condition where “{PATH_INFO} ‘does not match’ /site2”. That allows www.site2.com/site2 and www.site2.com to both function the same. This allows the site to always work, but if you want to hide /site2 in the URL, you need to take it one step further. One way to address this is in your code. Ultimately this is the best bet. Ruslan Yakushev has a great article on a few considerations that you can address in code. I recommend giving that serious consideration. Additionally, if you have upgraded to ASP.NET 3.5 SP1 or greater, it takes care of some of the references automatically for you. However, what if you inherit an existing application? Or you can’t easily go through your existing site and make the code changes? If this applies to you, read on. That’s where URL Rewrite 2.0 comes in. With URL Rewrite 2.0, you can create an outgoing rule that will remove the /site2 before the page is sent back to the user. This means that you can take an existing application, host it in a subfolder of your site, and ensure that the URL never reveals that it’s in a subfolder. Performance Considerations Performance overhead is something to be mindful of. These outbound rules aren’t simply changing the server variables. The first rule I’ll cover below needs to parse the HTML body and pull out the path (i.e. /site2) on the way through. This will add overhead, possibly significant if you have large pages and a busy site. In other words, your mileage may vary and you may need to test to see the impact that these rules have. Don’t worry too much though. For many sites, the performance impact is negligible. So, how do we do it? Creating the Outgoing Rule There are really two things to keep in mind. First, ASP.NET applications frequently generate a URL that adds the /site2 back into the URL. In addition to URLs, they can be in form elements, img elements and the like. The goal is to find all of those situations and rewrite it on the way out. Let’s call this the ‘URL problem’. Second, and similarly, ASP.NET can send a LOCATION redirect that causes a redirect back to another page. Again, ASP.NET isn’t aware of the different URL and it will add the /site2 to the redirect. Form Authentication is a good example on when this occurs. Try to password protect a site running from a subfolder using forms auth and you’ll quickly find that the URL becomes www.site2.com/site2 again. Let’s term this the ‘redirect problem’. Solving the URL Problem – Outgoing Rule #1 Let’s create a rule that removes the /site2 from any URL. We want to remove it from relative URLs like /site2/something, or absolute URLs like http://www.site2.com/site2/something. Most URLs that ASP.NET creates will be relative URLs, but I figure that there may be some applications that piece together a full URL, so we might as well expect that situation. Let’s get started. First, create a new outbound rule. You can create the rule within the /site2 folder which will reduce the performance impact of the rule. Just a reminder that incoming rules for this situation won’t work in a subfolder … but outgoing rules will. Give it a name that makes sense to you, for example “Outgoing – URL paths”. Precondition. If you place the rule in the subfolder, it will only run for that site and folder, so there isn’t need for a precondition. Run it for all requests. If you place it in the root of the site, you may want to create a precondition for HTTP_HOST = ^(www\.)?site2\.com$. For the Match section, there are a few things to consider. For performance reasons, it’s best to match the least amount of elements that you need to accomplish the task. For my test cases, I just needed to rewrite the <a /> tag, but you may need to rewrite any number of HTML elements. Note that as long as you have the exclude /site2 rule in your incoming rule as I described in Part I, some elements that don’t show their URL—like your images—will work without removing the /site2 from them. That reduces the processing needed for this rule. Leave the “matching scope” at “Response” and choose the elements that you want to change. Set the pattern to “^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*)”. Make sure to replace ‘site2’ with your subfolder name in both places. Yes, I realize this is a pretty messy looking rule, but it handles a few situations. This rule will handle the following situations correctly: Original Rewritten using {R:1}{R:2} http://www.site2.com/site2/default.aspx http://www.site2.com/default.aspx http://www.site2.com/folder1/site2/default.aspx Won’t rewrite since it’s a sub-sub folder /site2/default.aspx /default.aspx site2/default.aspx /default.aspx /folder1/site2/default.aspx Won’t rewrite since it’s a sub-sub folder. For the conditions section, you can leave that be. Finally, for the rule, set the Action Type to “Rewrite” and set the Value to “{R:1}{R:2}”. The {R:1} and {R:2} are back references to the sections within parentheses. In other words, in http://domain.com/site2/something, {R:1} will be http://domain.com and {R:2} will be /something. If you view your rule from your web.config file (or applicationHost.config if it’s a global rule), it should look like this: <rule name="Outgoing - URL paths" enabled="true"> <match filterByTags="A" pattern="^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> Solving the Redirect Problem Outgoing Rule #2 The second issue that we can run into is with a client-side redirect. This is triggered by a LOCATION response header that is sent to the client. Forms authentication is a common example. To reproduce this, password protect your subfolder and watch how it redirects and adds the subfolder path back in. Notice in my test case the extra paths: http://site2.com/site2/login.aspx?ReturnUrl=%2fsite2%2fdefault.aspx I want to remove /site2 from both the URL and the ReturnUrl querystring value. For semi-readability, let’s do this in 2 separate rules, one for the URL and one for the querystring. Create a second rule. As with the previous rule, it can be created in the /site2 subfolder. In the URL Rewrite wizard, select Outbound rules –> “Blank Rule”. Fill in the following information: Name response_location URL Precondition Don’t set Match: Matching Scope Server Variable Match: Variable Name RESPONSE_LOCATION Match: Pattern ^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*) Conditions Don’t set Action Type Rewrite Action Properties {R:1}{R:2} It should end up like so: <rule name="response_location URL"> <match serverVariable="RESPONSE_LOCATION" pattern="^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> Outgoing Rule #3 Outgoing Rule #2 only takes care of the URL path, and not the querystring path. Let’s create one final rule to take care of the path in the querystring to ensure that ReturnUrl=%2fsite2%2fdefault.aspx gets rewritten to ReturnUrl=%2fdefault.aspx. The %2f is the HTML encoding for forward slash (/). Create a rule like the previous one, but with the following settings: Name response_location querystring Precondition Don’t set Match: Matching Scope Server Variable Match: Variable Name RESPONSE_LOCATION Match: Pattern (.*)%2fsite2(.*) Conditions Don’t set Action Type Rewrite Action Properties {R:1}{R:2} The config should look like this: <rule name="response_location querystring"> <match serverVariable="RESPONSE_LOCATION" pattern="(.*)%2fsite2(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> It’s possible to squeeze the last two rules into one, but it gets kind of confusing so I felt that it’s better to show it as two separate rules. Summary With the rules covered in these two parts, we’re able to have a site in a subfolder and make it appear as if it’s in the root of the site. Not only that, we can overcome automatic redirecting that is caused by ASP.NET, other scripting technologies, and especially existing applications. Following is an example of the incoming and outgoing rules necessary for a site called www.site2.com hosted in a subfolder called /site2. Remember that the outgoing rules can be placed in the /site2 folder instead of the in the root of the site. <rewrite> <rules> <rule name="site2.com in a subfolder" enabled="true" stopProcessing="true"> <match url=".*" /> <conditions logicalGrouping="MatchAll" trackAllCaptures="false"> <add input="{HTTP_HOST}" pattern="^(www\.)?site2\.com$" /> <add input="{PATH_INFO}" pattern="^/site2($|/)" negate="true" /> </conditions> <action type="Rewrite" url="/site2/{R:0}" /> </rule> </rules> <outboundRules> <rule name="Outgoing - URL paths" enabled="true"> <match filterByTags="A" pattern="^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> <rule name="response_location URL"> <match serverVariable="RESPONSE_LOCATION" pattern="^(?:site2|(.*//[_a-zA-Z0-9-\.]*)?/site2)(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> <rule name="response_location querystring"> <match serverVariable="RESPONSE_LOCATION" pattern="(.*)%2fsite2(.*)" /> <action type="Rewrite" value="{R:1}{R:2}" /> </rule> </outboundRules> </rewrite> If you run into any situations that aren’t caught by these rules, please let me know so I can update this to be as complete as possible. Happy URL Rewriting!

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Adding Client Validation To DataAnnotations DataType Attribute

- by srkirkland

The System.ComponentModel.DataAnnotations namespace contains a validation attribute called DataTypeAttribute, which takes an enum specifying what data type the given property conforms to. Here are a few quick examples: public class DataTypeEntity { [DataType(DataType.Date)] public DateTime DateTime { get; set; } [DataType(DataType.EmailAddress)] public string EmailAddress { get; set; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } This attribute comes in handy when using ASP.NET MVC, because the type you specify will determine what “template” MVC uses. Thus, for the DateTime property if you create a partial in Views/[loc]/EditorTemplates/Date.ascx (or cshtml for razor), that view will be used to render the property when using any of the Html.EditorFor() methods. One thing that the DataType() validation attribute does not do is any actual validation. To see this, let’s take a look at the EmailAddress property above. It turns out that regardless of the value you provide, the entity will be considered valid: //valid new DataTypeEntity {EmailAddress = "Foo"}; .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Hmmm. Since DataType() doesn’t validate, that leaves us with two options: (1) Create our own attributes for each datatype to validate, like [Date], or (2) add validation into the DataType attribute directly. In this post, I will show you how to hookup client-side validation to the existing DataType() attribute for a desired type. From there adding server-side validation would be a breeze and even writing a custom validation attribute would be simple (more on that in future posts). Validation All The Way Down Our goal will be to leave our DataTypeEntity class (from above) untouched, requiring no reference to System.Web.Mvc. Then we will make an ASP.NET MVC project that allows us to create a new DataTypeEntity and hookup automatic client-side date validation using the suggested “out-of-the-box” jquery.validate bits that are included with ASP.NET MVC 3. For simplicity I’m going to focus on the only DateTime field, but the concept is generally the same for any other DataType. Building a DataTypeAttribute Adapter To start we will need to build a new validation adapter that we can register using ASP.NET MVC’s DataAnnotationsModelValidatorProvider.RegisterAdapter() method. This method takes two Type parameters; The first is the attribute we are looking to validate with and the second is an adapter that should subclass System.Web.Mvc.ModelValidator. Since we are extending DataAnnotations we can use the subclass of ModelValidator called DataAnnotationsModelValidator<>. This takes a generic argument of type DataAnnotations.ValidationAttribute, which lucky for us means the DataTypeAttribute will fit in nicely. So starting from there and implementing the required constructor, we get: public class DataTypeAttributeAdapter : DataAnnotationsModelValidator<DataTypeAttribute> { public DataTypeAttributeAdapter(ModelMetadata metadata, ControllerContext context, DataTypeAttribute attribute) : base(metadata, context, attribute) { } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now you have a full-fledged validation adapter, although it doesn’t do anything yet. There are two methods you can override to add functionality, IEnumerable<ModelValidationResult> Validate(object container) and IEnumerable<ModelClientValidationRule> GetClientValidationRules(). Adding logic to the server-side Validate() method is pretty straightforward, and for this post I’m going to focus on GetClientValidationRules(). Adding a Client Validation Rule Adding client validation is now incredibly easy because jquery.validate is very powerful and already comes with a ton of validators (including date and regular expressions for our email example). Teamed with the new unobtrusive validation javascript support we can make short work of our ModelClientValidationDateRule: public class ModelClientValidationDateRule : ModelClientValidationRule { public ModelClientValidationDateRule(string errorMessage) { ErrorMessage = errorMessage; ValidationType = "date"; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } If your validation has additional parameters you can the ValidationParameters IDictionary<string,object> to include them. There is a little bit of conventions magic going on here, but the distilled version is that we are defining a “date” validation type, which will be included as html5 data-* attributes (specifically data-val-date). Then jquery.validate.unobtrusive takes this attribute and basically passes it along to jquery.validate, which knows how to handle date validation. Finishing our DataTypeAttribute Adapter Now that we have a model client validation rule, we can return it in the GetClientValidationRules() method of our DataTypeAttributeAdapter created above. Basically I want to say if DataType.Date was provided, then return the date rule with a given error message (using ValidationAttribute.FormatErrorMessage()). The entire adapter is below: public class DataTypeAttributeAdapter : DataAnnotationsModelValidator<DataTypeAttribute> { public DataTypeAttributeAdapter(ModelMetadata metadata, ControllerContext context, DataTypeAttribute attribute) : base(metadata, context, attribute) { } public override System.Collections.Generic.IEnumerable<ModelClientValidationRule> GetClientValidationRules() { if (Attribute.DataType == DataType.Date) { return new[] { new ModelClientValidationDateRule(Attribute.FormatErrorMessage(Metadata.GetDisplayName())) }; } return base.GetClientValidationRules(); } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Putting it all together Now that we have an adapter for the DataTypeAttribute, we just need to tell ASP.NET MVC to use it. The easiest way to do this is to use the built in DataAnnotationsModelValidatorProvider by calling RegisterAdapter() in your global.asax startup method. DataAnnotationsModelValidatorProvider.RegisterAdapter(typeof(DataTypeAttribute), typeof(DataTypeAttributeAdapter)); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Show and Tell Let’s see this in action using a clean ASP.NET MVC 3 project. First make sure to reference the jquery, jquery.vaidate and jquery.validate.unobtrusive scripts that you will need for client validation. Next, let’s make a model class (note we are using the same built-in DataType() attribute that comes with System.ComponentModel.DataAnnotations). public class DataTypeEntity { [DataType(DataType.Date, ErrorMessage = "Please enter a valid date (ex: 2/14/2011)")] public DateTime DateTime { get; set; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Then we make a create page with a strongly-typed DataTypeEntity model, the form section is shown below (notice we are just using EditorForModel): @using (Html.BeginForm()) { @Html.ValidationSummary(true) <fieldset> <legend>Fields</legend> @Html.EditorForModel() <p> <input type="submit" value="Create" /> </p> </fieldset> } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The final step is to register the adapter in our global.asax file: DataAnnotationsModelValidatorProvider.RegisterAdapter(typeof(DataTypeAttribute), typeof(DataTypeAttributeAdapter)); Now we are ready to run the page: Looking at the datetime field’s html, we see that our adapter added some data-* validation attributes: <input type="text" value="1/1/0001" name="DateTime" id="DateTime" data-val-required="The DateTime field is required." data-val-date="Please enter a valid date (ex: 2/14/2011)" data-val="true" class="text-box single-line valid"> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Here data-val-required was added automatically because DateTime is non-nullable, and data-val-date was added by our validation adapter. Now if we try to add an invalid date: Our custom error message is displayed via client-side validation as soon as we tab out of the box. If we didn’t include a custom validation message, the default DataTypeAttribute “The field {0} is invalid” would have been shown (of course we can change the default as well). Note we did not specify server-side validation, but in this case we don’t have to because an invalid date will cause a server-side error during model binding. Conclusion I really like how easy it is to register new data annotations model validators, whether they are your own or, as in this post, supplements to existing validation attributes. I’m still debating about whether adding the validation directly in the DataType attribute is the correct place to put it versus creating a dedicated “Date” validation attribute, but it’s nice to know either option is available and, as we’ve seen, simple to implement. I’m also working through the nascent stages of an open source project that will create validation attribute extensions to the existing data annotations providers using similar techniques as seen above (examples: Email, Url, EqualTo, Min, Max, CreditCard, etc). Keep an eye on this blog and subscribe to my twitter feed (@srkirkland) if you are interested for announcements.

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Can't connect to WCF service on Android

- by illvm

I am trying to connect to a .NET WCF service in Android using kSOAP2 (v 2.1.2), but I keep getting a fatal exception whenever I try to make the service call. I'm having a bit of difficulty tracking down the error and can't seem to figure out why it's happening. The code I am using is below: package org.example.android; import org.ksoap2.SoapEnvelope; import org.ksoap2.serialization.PropertyInfo; import org.ksoap2.serialization.SoapObject; import org.ksoap2.serialization.SoapSerializationEnvelope; import org.ksoap2.transport.HttpTransport; import android.app.Activity; import android.app.AlertDialog; import android.content.DialogInterface; import android.content.Intent; import android.os.Bundle; public class ValidateUser extends Activity { private static final String SOAP_ACTION = "http://tempuri.org/mobile/ValidateUser"; private static final String METHOD_NAME = "ValidateUser"; private static final String NAMESPACE = "http://tempuri.org/mobile/"; private static final String URL = "http://192.168.1.2:8002/WebService.Mobile.svc"; /** Called when the activity is first created. */ @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.validate_user); Intent intent = getIntent(); Bundle extras = intent.getExtras(); if (extras == null) { this.finish(); } String username = extras.getString("username"); String password = extras.getString("password"); Boolean validUser = false; try { SoapObject request = new SoapObject(NAMESPACE, METHOD_NAME); PropertyInfo uName = new PropertyInfo(); uName.name = "userName"; PropertyInfo pWord = new PropertyInfo(); pWord.name = "passWord"; request.addProperty(uName, username); request.addProperty(pWord, password); SoapSerializationEnvelope envelope = new SoapSerializationEnvelope(SoapEnvelope.VER11); envelope.dotNet = true; envelope.setOutputSoapObject(request); HttpTransport androidHttpTransport = new HttpTransport(URL); androidHttpTransport.call(SOAP_ACTION, envelope); // error occurs here Integer userId = (Integer)envelope.getResponse(); validUser = (userId != 0); } catch (Exception ex) { } } private void exit () { this.finish(); } } EDIT: Remove the old stack traces. In summary, the first problem was the program being unable to open a connection due to missing methods or libraries due to using vanilla kSOAP2 rather than a modified library for Android (kSOAP2-Android). The second issue was a settings issue. In the Manifest I did not add the following setting: <uses-permission android:name="android.permission.INTERNET" /> I am now having an issue with the XMLPullParser which I need to figure out. 12-23 10:58:06.480: ERROR/SOCKETLOG(210): add_recv_stats recv 0 12-23 10:58:06.710: WARN/System.err(210): org.xmlpull.v1.XmlPullParserException: unexpected type (position:END_DOCUMENT null@1:0 in java.io.InputStreamReader@433fb070) 12-23 10:58:06.710: WARN/System.err(210): at org.kxml2.io.KXmlParser.exception(KXmlParser.java:243) 12-23 10:58:06.720: WARN/System.err(210): at org.kxml2.io.KXmlParser.nextTag(KXmlParser.java:1363) 12-23 10:58:06.720: WARN/System.err(210): at org.ksoap2.SoapEnvelope.parse(SoapEnvelope.java:126) 12-23 10:58:06.720: WARN/System.err(210): at org.ksoap2.transport.Transport.parseResponse(Transport.java:63) 12-23 10:58:06.720: WARN/System.err(210): at org.ksoap2.transport.HttpTransportSE.call(HttpTransportSE.java:100) 12-23 10:58:06.730: WARN/System.err(210): at org.example.android.ValidateUser.onCreate(ValidateUser.java:68) 12-23 10:58:06.730: WARN/System.err(210): at android.app.Instrumentation.callActivityOnCreate(Instrumentation.java:1122) 12-23 10:58:06.730: WARN/System.err(210): at android.app.ActivityThread.performLaunchActivity(ActivityThread.java:2104) 12-23 10:58:06.730: WARN/System.err(210): at android.app.ActivityThread.handleLaunchActivity(ActivityThread.java:2157) 12-23 10:58:06.730: WARN/System.err(210): at android.app.ActivityThread.access$1800(ActivityThread.java:112) 12-23 10:58:06.730: WARN/System.err(210): at android.app.ActivityThread$H.handleMessage(ActivityThread.java:1581) 12-23 10:58:06.730: WARN/System.err(210): at android.os.Handler.dispatchMessage(Handler.java:88) 12-23 10:58:06.730: WARN/System.err(210): at android.os.Looper.loop(Looper.java:123) 12-23 10:58:06.730: WARN/System.err(210): at android.app.ActivityThread.main(ActivityThread.java:3739) 12-23 10:58:06.730: WARN/System.err(210): at java.lang.reflect.Method.invokeNative(Native Method) 12-23 10:58:06.730: WARN/System.err(210): at java.lang.reflect.Method.invoke(Method.java:515) 12-23 10:58:06.730: WARN/System.err(210): at com.android.internal.os.ZygoteInit$MethodAndArgsCaller.run(ZygoteInit.java:739) 12-23 10:58:06.730: WARN/System.err(210): at com.android.internal.os.ZygoteInit.main(ZygoteInit.java:497) 12-23 10:58:06.730: WARN/System.err(210): at dalvik.system.NativeStart.main(Native Method)

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Jetty 7 + MySQL Config [java.lang.ClassNotFoundException: org.mortbay.jetty.webapp.WebAppContext]

- by Scott Chang

I've been trying to get a c3p0 db connection pool configured for Jetty, but I keep getting a ClassNotFoundException: 2010-03-14 19:32:12.028:WARN::Failed startup of context WebAppContext@fccada@fccada/phpMyAdmin,file:/usr/local/jetty/webapps/phpMyAdmin/,file:/usr/local/jetty/webapps/phpMyAdmin/ java.lang.ClassNotFoundException: org.mortbay.jetty.webapp.WebAppContext at java.net.URLClassLoader$1.run(URLClassLoader.java:200) at java.security.AccessController.doPrivileged(Native Method) at java.net.URLClassLoader.findClass(URLClassLoader.java:188) at java.lang.ClassLoader.loadClass(ClassLoader.java:307) at java.lang.ClassLoader.loadClass(ClassLoader.java:252) at org.eclipse.jetty.webapp.WebAppClassLoader.loadClass(WebAppClassLoader.java:313) at org.eclipse.jetty.webapp.WebAppClassLoader.loadClass(WebAppClassLoader.java:266) at org.eclipse.jetty.util.Loader.loadClass(Loader.java:90) at org.eclipse.jetty.xml.XmlConfiguration.nodeClass(XmlConfiguration.java:224) at org.eclipse.jetty.xml.XmlConfiguration.configure(XmlConfiguration.java:187) at org.eclipse.jetty.webapp.JettyWebXmlConfiguration.configure(JettyWebXmlConfiguration.java:77) at org.eclipse.jetty.webapp.WebAppContext.startContext(WebAppContext.java:975) at org.eclipse.jetty.server.handler.ContextHandler.doStart(ContextHandler.java:586) at org.eclipse.jetty.webapp.WebAppContext.doStart(WebAppContext.java:349) at org.eclipse.jetty.util.component.AbstractLifeCycle.start(AbstractLifeCycle.java:55) at org.eclipse.jetty.server.handler.HandlerCollection.doStart(HandlerCollection.java:165) at org.eclipse.jetty.server.handler.ContextHandlerCollection.doStart(ContextHandlerCollection.java:162) at org.eclipse.jetty.util.component.AbstractLifeCycle.start(AbstractLifeCycle.java:55) at org.eclipse.jetty.server.handler.HandlerCollection.doStart(HandlerCollection.java:165) at org.eclipse.jetty.util.component.AbstractLifeCycle.start(AbstractLifeCycle.java:55) at org.eclipse.jetty.server.handler.HandlerWrapper.doStart(HandlerWrapper.java:92) at org.eclipse.jetty.server.Server.doStart(Server.java:228) at org.eclipse.jetty.util.component.AbstractLifeCycle.start(AbstractLifeCycle.java:55) at org.eclipse.jetty.xml.XmlConfiguration$1.run(XmlConfiguration.java:990) at java.security.AccessController.doPrivileged(Native Method) at org.eclipse.jetty.xml.XmlConfiguration.main(XmlConfiguration.java:955) at sun.reflect.NativeMethodAccessorImpl.invoke0(Native Method) at sun.reflect.NativeMethodAccessorImpl.invoke(NativeMethodAccessorImpl.java:39) at sun.reflect.DelegatingMethodAccessorImpl.invoke(DelegatingMethodAccessorImpl.java:25) at java.lang.reflect.Method.invoke(Method.java:597) at org.eclipse.jetty.start.Main.invokeMain(Main.java:394) at org.eclipse.jetty.start.Main.start(Main.java:546) at org.eclipse.jetty.start.Main.parseCommandLine(Main.java:208) at org.eclipse.jetty.start.Main.main(Main.java:75) I'm new to Jetty and I want to ultimately get phpMyAdmin and WordPress to run on it through Quercus and a JDBC connection. Here are my web.xml and jetty-web.xml files in my WEB-INF directory. jetty-web.xml: <?xml version="1.0"?> <!DOCTYPE Configure PUBLIC "-//Mort Bay Consulting//DTD Configure//EN" "http://jetty.mortbay.org/configure.dtd"> <Configure class="org.mortbay.jetty.webapp.WebAppContext"> <New id="mysql" class="org.mortbay.jetty.plus.naming.Resource"> <Arg>jdbc/mysql</Arg> <Arg> <New class="com.mchange.v2.c3p0.ComboPooledDataSource"> <Set name="Url">jdbc:mysql://localhost:3306/mysql</Set> <Set name="User">user</Set> <Set name="Password">pw</Set> </New> </Arg> </New> </Configure> web.xml: <?xml version="1.0"?> <!DOCTYPE web-app PUBLIC "-//Sun Microsystems, Inc.//DTD Web Application 2.2//EN" "http://java.sun.com/j2ee/dtds/web-app_2_2.dtd"> <web-app> <description>Caucho Technology's PHP Implementation</description> <resource-ref> <description>My DataSource Reference</description> <res-ref-name>jdbc/mysql</res-ref-name> <res-type>javax.sql.DataSource</res-type> <res-auth>Container</res-auth> </resource-ref> <servlet> <servlet-name>Quercus Servlet</servlet-name> <servlet-class>com.caucho.quercus.servlet.QuercusServlet</servlet-class>  <init-param> <param-name>script-encoding</param-name> <param-value>UTF-8</param-value> </init-param>  <init-param> <param-name>database</param-name> <param-value>jdbc/mysql</param-value> </init-param> <init-param> <param-name>ini-file</param-name> <param-value>WEB-INF/php.ini</param-value> </init-param> </servlet> <servlet-mapping> <servlet-name>Quercus Servlet</servlet-name> <url-pattern>*.php</url-pattern> </servlet-mapping> <welcome-file-list> <welcome-file>index.php</welcome-file> </welcome-file-list> </web-app> I'm guessing that I'm missing a few jars or something. Currently I have placed the following jars in my WEB-INF/lib directory: c3p0-0.9.1.2.jar commons-dbcp-1.4.jar commons-pool-1.5.4.jar mysql-connector-java-5.1.12-bin.jar I have also tried to put these jars in JETTY-HOME/lib/ext, but to no avail... Someone please tell me what is wrong with my configuration. I'm sick of digging through Jetty's crappy documentation.

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SUPER CSV write bean to CSV.

- by ButtersB

Here is my class, public class FreebasePeopleResults { public String intendedSearch; public String weight; public Double heightMeters; public Integer age; public String type; public String parents; public String profession; public String alias; public String children; public String siblings; public String spouse; public String degree; public String institution; public String wikipediaId; public String guid; public String id; public String gender; public String name; public String ethnicity; public String articleText; public String dob; public String getWeight() { return weight; } public void setWeight(String weight) { this.weight = weight; } public Double getHeightMeters() { return heightMeters; } public void setHeightMeters(Double heightMeters) { this.heightMeters = heightMeters; } public String getParents() { return parents; } public void setParents(String parents) { this.parents = parents; } public Integer getAge() { return age; } public void setAge(Integer age) { this.age = age; } public String getProfession() { return profession; } public void setProfession(String profession) { this.profession = profession; } public String getAlias() { return alias; } public void setAlias(String alias) { this.alias = alias; } public String getChildren() { return children; } public void setChildren(String children) { this.children = children; } public String getSpouse() { return spouse; } public void setSpouse(String spouse) { this.spouse = spouse; } public String getDegree() { return degree; } public void setDegree(String degree) { this.degree = degree; } public String getInstitution() { return institution; } public void setInstitution(String institution) { this.institution = institution; } public String getWikipediaId() { return wikipediaId; } public void setWikipediaId(String wikipediaId) { this.wikipediaId = wikipediaId; } public String getGuid() { return guid; } public void setGuid(String guid) { this.guid = guid; } public String getId() { return id; } public void setId(String id) { this.id = id; } public String getGender() { return gender; } public void setGender(String gender) { this.gender = gender; } public String getName() { return name; } public void setName(String name) { this.name = name; } public String getEthnicity() { return ethnicity; } public void setEthnicity(String ethnicity) { this.ethnicity = ethnicity; } public String getArticleText() { return articleText; } public void setArticleText(String articleText) { this.articleText = articleText; } public String getDob() { return dob; } public void setDob(String dob) { this.dob = dob; } public String getType() { return type; } public void setType(String type) { this.type = type; } public String getSiblings() { return siblings; } public void setSiblings(String siblings) { this.siblings = siblings; } public String getIntendedSearch() { return intendedSearch; } public void setIntendedSearch(String intendedSearch) { this.intendedSearch = intendedSearch; } } Here is my CSV writer method import java.io.FileWriter; import java.io.IOException; import java.util.ArrayList; import org.supercsv.io.CsvBeanWriter; import org.supercsv.prefs.CsvPreference; public class CSVUtils { public static void writeCSVFromList(ArrayList<FreebasePeopleResults> people, boolean writeHeader) throws IOException{ //String[] header = new String []{"title","acronym","globalId","interfaceId","developer","description","publisher","genre","subGenre","platform","esrb","reviewScore","releaseDate","price","cheatArticleId"}; FileWriter file = new FileWriter("/brian/brian/Documents/people-freebase.csv", true); // write the partial data CsvBeanWriter writer = new CsvBeanWriter(file, CsvPreference.EXCEL_PREFERENCE); for(FreebasePeopleResults person:people){ writer.write(person); } writer.close(); // show output } } I keep getting output errors. Here is the error: There is no content to write for line 2 context: Line: 2 Column: 0 Raw line: null Now, I know it is now totally null, so I am confused.

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Jolicloud is a Nifty New OS for Your Netbook

- by Matthew Guay

Want to breathe new life into your netbook? Here’s a quick look at Jolicloud, a unique new Linux based OS that lets you use your netbook in a whole new way. Netbooks have been an interesting category of computers. When they were first released, most netbooks came with a stripped down Linux based operating system designed to let you easily access the internet first and foremost. Consumers wanted more from their netbooks, so full OSes such as Windows XP and Ubuntu became the standard on netbooks. Microsoft worked hard to get Windows 7 working great on netbooks, and today most netbooks run Windows 7 great. But the Linux community hasn’t stood still either, and Jolicloud is proof of that. Jolicloud is a unique OS designed to bring the best of both webapps and standard programs to your netbook. Keep reading to see if this is the perfect netbook OS for you. Getting Started Installing Jolicloud on your netbook is easy thanks to a the Jolicloud Express installer for Windows. Since many netbooks run Windows by default, this makes it easy to install Jolicloud. Plus, your Windows install is left untouched, so you can still easily access all your Windows files and programs. Download and run the roughly 700Mb installer (link below) just as a normal installer in Windows. This will first extract the needed files. Click Get started to install Jolicloud on your netbook. Enter a username, password, and nickname for your computer. Please note that the username must be all lowercase, and the nickname should not contain spaces or special characters. Now you can review the default installation settings. By default it will take up 39Gb and install on your C:\ drive in English. If you wish to change this, click Change. We chose to install it on the D: drive on this netbook, as its harddrive was already partitioned into two parts. Click Save when your settings are all correct, and then click Next in the previous window. Jolicloud will prepare for the installation. This took about 5 minutes in our test. Click Next when this is finished. Click Restart now to install and run Jolicloud. When your netbook reboots, it will initialize the Jolicloud setup. It will then automatically finish the installation. Just sit back and wait; there’s nothing for you to do right now. The installation took about 20 minutes in our test. Jolicloud will automatically reboot when the setup is finished. Once it’s rebooted, you’re ready to go! Enter the username, then the password, that you chose earlier when you were installing Jolicloud from Windows. Welcome to your Jolicloud desktop! Hardware Support We installed Jolicloud on a Samsung N150 netbook with an Atom N450 processor, 1Gb Ram, 250Gb harddrive, and WiFi b/g/n with Bluetooth. Amazingly, once Jolicloud was installed, everything was ready to use. No drivers to install, no settings to hassle with, it was all installed and set up perfectly. Power settings worked great, and closing the netbook put it to sleep just like in Windows. WiFi drivers have typically been difficult to find and install on Linux, but Jolicloud had our netbook’s wifi working immediately. To get online, simply click the Wireless icon on the top right, and select the wireless network you want to connect to. Jolicloud will let you know when it is signed on. Wired Lan networking was also seamless; simply connect your cable and you’re ready to go. The webcam and touchpad also worked perfectly directly. The only thing missing was multitouch; this touchpad has two finger scroll, pinch zoom, and other nice multitouch features in Windows, but in Julicloud it only functioned as a standard touchpad. It did have tap to click activated by default, as well as right-side scrolling, which is nice. Jolicloud also supported our video card without any extra work. The native resolution was already selected, and the only problem we had with the screen was that there was no apparent way to change the brightness. This is not a major problem, but would be nice to have. The Samsung N150 has Intel GMA3150 integrated graphics, and Jolicloud promises 1080p HD video on it. It did playback 720p H.264 video flawlessly without installing anything extra, but it stuttered on full 1080p HD (which is the exact same as this netbook’s video playback in Windows 7 – 720p works great, but it stutters on 1080p). We would be excited to see full HD on this netbook, but 720p is definitely fine for most stuff. Jolicloud supports a wide range of netbooks, and based on our experience we would expect it to work as good on any supported hardware. Check out the list of supported netbooks to see if your netbook is supported; if not, it still may work but you may have to install special drivers. Jolicloud’s performance was very similar to Windows 7 on our netbook. It boots in about 30 seconds, and apps load fairly quickly. In general, we couldn’t tell much difference in performance between Jolicloud and Windows 7, though this isn’t a problem since Windows 7 runs great on the current generation of netbooks. Using Jolicloud Ready to start putting Jolicloud to use? Your fresh Jolicloud install you can run several built-in apps, such as Firefox, a calculator, and the chat client Pidgin. It also has a media player and file viewer installed, so you can play MP3s or MPG videos, or read PDF ebooks without installing anything extra. It also has Flash player installed so you can watch videos online easily. You can also directly access all of your files from the right side of your home screen. You can even access your Windows files; in our test, the 116.9 GB Media was C: from Windows. Select it to browse and open any file you had saved in Windows. You may need to enter your password to access it. Once you’re authenticated it, you’ll see all of your Windows files and folders. Your User files (Documents, Music, Videos, etc.) will be in the Users folder. And, you can easily add files from removable media such as USB flash drives and memory cards. Jolicloud recognized a flash drive we tested with no trouble at all. Add new apps But, the best part about Jolicloud is that it makes it very easy to install new apps. Click the Get Started button on your homescreen. You’ll first need to create an account. You can then use this same account on another netbook if you wish, and your settings will automatically be synced between the two. You can either signup using your Facebook account, …or you can sign up the traditional way with your email address, name, and password. If you sign up this way, you will need to confirm your email address before your account will be finished. Now, choose your netbook model from the list, and enter a name for your computer. And that’s it! You’ll now see the Jolicloud dashboard, which will show you updates and notifications from friends who also use Jolicloud. Click the App directory to find new apps for your netbook. Here you will find a variety of webapps, such as Gmail, along with native applications, such as Skype, that you can install on your netbook. Simply click the Install button on the right to add the app to your netbook. You will be prompted to enter your system password, and then the app will install without any further input. Once an app is installed, a check mark will appear beside its name. You can remove it by clicking the Remove button, and it will uninstall seamlessly. Webapps, such as Gmail, actually run in in a Chrome-powered window that lets the webapp run full screen. This gives the webapps a native feel, but actually they’re just running the same as they would in a standard web browser. The Jolicloud Interface Most apps run maximized, and there is no way to run them smaller. This in general works good, since with small screens most apps need to run full-screen anyhow. Smaller apps, such as a calculator or the Pidgin chat client, run in a window just like they do on other operating systems. You can switch to another app that’s running by selecting it’s icon on the top left, or you can go back to the home screen by clicking the home screen. If you’re finished with an program, simply click the red X button on the top right of the window when you’re running it. Or, you can switch between programs using standard keyboard shortcuts such as Alt-tab. The default page on the home screen is the favorites page, and all of your other programs are orginized in their own sections on the left hand side. But, if you want to add one of these to your favorites page, simply right-click on it and select Add to Favorites. When you’re done for the day, you can simply close your netbook to put it to sleep. Or, if you want to shut down, just press the Quit button on the bottom right of the home screen and then select Shut Down. Booting Jolicloud When you install Jolicloud, it will set itself as the default operating system. Now, when you boot your netbook, it will show you a list of installed operating systems. You can select either Windows or Jolicloud, but if you don’t make a selection it will boot into Jolicloud after waiting 10 seconds. If you’d perfer to boot into Windows by default, you can easily change this. First, boot your netbook in to Windows. Open the start menu, right-click on the Computer button, and select Properties. Click the “Advanced system settings” link on the left side. Click the Settings button in the Startup and Recovery section. Now, select Windows as the default operating system, and click Ok. Your netbook will now boot into Windows by default, but will give you 10 seconds to choose to boot into Jolicloud when you start your computer. Or, if you decided you don’t want Jolicloud, you can easily uninstall it from within Windows. Please note that this will also remove any files you may have saved in Jolicloud, so be sure to copy them to your Windows drive before uninstalling. To uninstall Jolicloud from within Windows, open Control Panel, and select Uninstall a Program. Scroll down to select Jolicloud, and click Uninstall/Change. Click Yes to confirm that you want to uninstall Jolicloud. After a few moments, it will let you know that Jolicloud has been uninstalled. You’re netbook is now back the same as it was before you installed Jolicloud, with only Windows installed. Closing Whether you’re wanting to replace your current OS on your netbook or would simply like to try out a fresh new Linux version on your netbook, Jolicloud is a great option for you. We were very impressed by it’s solid hardware support and the ease of installing new apps in Jolicloud. Rather than simply giving us a standard OS, Jolicloud offers a unique way to use your netbook with native programs and webapps. And whether you’re an IT pro or are a new computer user, Jolicloud was easy enough to use that anyone can do it. Give it a try, and let us know what your favorite netbook OS is! Link Download Jolicloud for your netbook Similar Articles Productive Geek Tips How To Change XSplash Themes in Ubuntu 9.10Verify the Integrity of Windows Vista System FilesMonitor Multiple Logs in a Single Shell with MultiTail for LinuxHide Some or All of the GUI Bars in FirefoxAsk the Readers: Do You Use a Laptop, Desktop, or Both? 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Configure Forms based authentication in SharePoint 2010

- by sreejukg

Configuring form authentication is a straight forward task in SharePoint. Mostly public facing websites built on SharePoint requires form based authentication. Recently, one of the WCM implementation where I was included in the project team required registration system. Any internet user can register to the site and the site offering them some membership specific functionalities once the user logged in. Since the registration open for all, I don’t want to store all those users in Active Directory. I have decided to use Forms based authentication for those users. This is a typical scenario of form authentication in SharePoint implementation. To implement form authentication you require the following A data store where you are storing the users – technically this can be active directory, SQL server database, LDAP etc. Form authentication will redirect the user to the login page, if the request is not authenticated. In the login page, there will be controls that validate the user inputs against the configured data store. In this article, I am going to use SQL server database with ASP.Net membership API’s to configure form based authentication in SharePoint 2010. This article assumes that you have SQL membership database available. I already configured the membership and roles database using aspnet_regsql command. If you want to know how to configure membership database using aspnet_regsql command, read the below blog post. http://weblogs.asp.net/sreejukg/archive/2011/06/16/usage-of-aspnet-regsql-exe-in-asp-net-4.aspx The snapshot of the database after implementing membership and role manager is as follows. I have used the database name “aspnetdb_claim”. Make sure you have created the database and make sure your database contains tables and stored procedures for membership. Create a web application with claims based authentication. This article assumes you already created a web application using claims based authentication. If you want to enable forms based authentication in SharePoint 2010, you must enable claims based authentication. Read this post for creating a web application using claims based authentication. http://weblogs.asp.net/sreejukg/archive/2011/06/15/create-a-web-application-in-sharepoint-2010-using-claims-based-authentication.aspx You make sure, you have selected enable form authentication, and then selected Membership provider and Role manager name. To make sure you are done with the configuration, navigate to central administration website, from central administration, navigate to the Web Applications page, select the web application and click on icon, you will see the authentication providers for the current web application. Go to the section Claims authentication types, and make sure you have enabled forms based authentication. As mentioned in the snapshot, I have named the membership provider as SPFormAuthMembership and role manager as SPFormAuthRoleManager. You can choose your own names as you need. Modify the configuration files(Web.Config) to enable form authentication There are three applications that needs to be configured to support form authentication. The following are those applications. Central Administration If you want to assign permissions to web application using the credentials from form authentication, you need to update Central Administration configuration. If you do not want to access form authentication credentials from Central Administration, just leave this step. STS service application Security Token service is the service application that issues security token when users are logging in. You need to modify the configuration of STS application to make sure users are able to login. To find the STS application, follow the following steps Go to the IIS Manager Expand the sites Node, you will see SharePoint Web Services Expand SharePoint Web Services, you can see SecurityTokenServiceApplication Right click SecuritytokenServiceApplication and click explore, it will open the corresponding file system. By default, the path for STS is C:\Program Files\Common Files\Microsoft Shared\Web Server Extensions\14\WebServices\SecurityToken You need to modify the configuration file available in the mentioned location. The web application that needs to be enabled with form authentication. You need to modify the configuration of your web application to make sure your web application identifies users from the form authentication. Based on the above, I am going to modify the web configuration. At end of each step, I have mentioned the expected output. I recommend you to go step by step and after each step, make sure the configuration changes are working as expected. If you do everything all together, and test your application at the end, you may face difficulties in troubleshooting the configuration errors. Modifications for Central Administration Web.Config Open the web.config for the Central administration in a text editor. I always prefer Visual Studio, for editing web.config. In most cases, the path of the web.config for the central administration website is as follows C:\inetpub\wwwroot\wss\VirtualDirectories\<port number> Make sure you keep a backup copy of the web.config, before editing it. Let me summarize what we are going to do with Central Administration web.config. First I am going to add a connection string that points to the form authentication database, that I created as mentioned in previous steps. Then I need to add a membership provider and a role manager with the corresponding connectionstring. Then I need to update the peoplepickerwildcards section to make sure the users are appearing in search results. By default there is no connection string available in the web.config of Central Administration. Add a connection string just after the configsections element. The below is the connection string I have used all over the article. <add name="FormAuthConnString" connectionString="Initial Catalog=yourdatabasename;data source=databaseservername;Integrated Security=SSPI;" /> Once you added the connection string, the web.config look similar to Now add membership provider to the code. In web.config for CA, there will be <membership> tag, search for it. You will find membership and role manager under the <system.web> element. Under the membership providers section add the below code… <add name="SPFormAuthMembership" type="System.Web.Security.SqlMembershipProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" applicationName="FormAuthApplication" connectionStringName="FormAuthConnString" /> After adding memberhip element, see the snapshot of the web.config. Now you need to add role manager element to the web.config. Insider providers element under rolemanager, add the below code. <add name="SPFormAuthRoleManager" type="System.Web.Security.SqlRoleProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" applicationName="FormAuthApplication" connectionStringName="FormAuthConnString" /> After adding, your role manager will look similar to the following. As a last step, you need to update the people picker wildcard element in web.config, so that the users from your membership provider are available for browsing in Central Administration. Search for PeoplePickerWildcards in the web.config, add the following inside the <PeoplePickerWildcards> tag. <add key="SPFormAuthMembership" value="%" /> After adding this element, your web.config will look like After completing these steps, you can browse the users available in the SQL server database from central administration website. Go to the site collection administrator’s page from central administration. Select the site collection you have created for form authentication. Click on the people picker icon, choose Forms Auth and click on the search icon, you will see the users listed from the SQL server database. Once you complete these steps, make sure the users are available for browsing from central administration website. If you are unable to find the users, there must be some errors in the configuration, check windows event logs to find related errors and fix them. Change the web.config for STS application Open the web.config for STS application in text editor. By default, STS web.config does not have system.Web or connectionstrings section. Just after the System.Webserver element, add the following code. <connectionStrings> <add name="FormAuthConnString" connectionString="Initial Catalog=aspnetdb_claim;data source=sp2010_db;Integrated Security=SSPI;" /> </connectionStrings> <system.web> <roleManager enabled="true" cacheRolesInCookie="false" cookieName=".ASPXROLES" cookieTimeout="30" cookiePath="/" cookieRequireSSL="false" cookieSlidingExpiration="true" cookieProtection="All" createPersistentCookie="false" maxCachedResults="25"> <providers> <add name="SPFormAuthRoleManager" type="System.Web.Security.SqlRoleProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" applicationName="FormAuthApplication" connectionStringName="FormAuthConnString" /> </providers> </roleManager> <membership userIsOnlineTimeWindow="15" hashAlgorithmType=""> <providers> <add name="SPFormAuthMembership" type="System.Web.Security.SqlMembershipProvider, System.Web, Version=2.0.0.0, Culture=neutral, PublicKeyToken=b03f5f7f11d50a3a" applicationName="FormAuthApplication" connectionStringName="FormAuthConnString" /> </providers> </membership> </system.web> See the snapshot of the web.config after adding the required elements. After adding this, you should be able to login using the credentials from SQL server. Try assigning a user as primary/secondary administrator for your site collection from Central Administration and login to your site using form authentication. If you made everything correct, you should be able to login. This means you have successfully completed configuration of STS Configuration of Web Application for Form Authentication As a last step, you need to modify the web.config of the form authentication web application. Once you have done this, you should be able to grant permissions to users stored in the membership database. Open the Web.config of the web application you created for form authentication. You can find the web.config for the application under the path C:\inetpub\wwwroot\wss\VirtualDirectories\<port number> Basically you need to add connection string, membership provider, role manager and update the people picker wild card configuration. Add the connection string (same as the one you added to the web.config in Central Administration). See the screenshot after the connection string has added. Search for <membership> in the web.config, you will find this inside system.web element. There will be other providers already available there. You add your form authentication membership provider (similar to the one added to Central Administration web.config) to the provider element under membership. Find the snapshot of membership configuration as follows. Search for <roleManager> element in web.config, add the new provider name under providers section of the roleManager element. See the snapshot of web.config after new provider added. Now you need to configure the peoplepickerwildcard configuration in web.config. As I specified earlier, this is to make sure, you can locate the users by entering a part of their username. Add the following line under the <PeoplePickerWildcards> element in web.config. See the screenshot of the peoplePickerWildcards element after the element has been added. Now you have completed all the setup for form authentication. Navigate to the web application. From the site actions -> site settings -> go to peope and groups Click on new -> add users, it will popup the people picker dialog. Click on the icon, select Form Auth, enter a username in the search textbox, and click on search icon. See the screenshot of admin search when I tried searching the users If it displays the user, it means you are done with the configuration. If you add users to the form authentication database, the users will be able to access SharePoint portal as normal.

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