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  • How to convert objects in reflection (Linq to XML)

    - by user829174
    I have the following code which is working fine, the code creates plugins in reflection via run time: using System; using System.Collections.Generic; using System.Linq; using System.IO; using System.Xml.Linq; using System.Reflection; using System.Text; namespace WindowsFormsApplication1 { public abstract class Plugin { public string Type { get; set; } public string Message { get; set; } } public class FilePlugin : Plugin { public string Path { get; set; } } public class RegsitryPlugin : Plugin { public string Key { get; set; } public string Name { get; set; } public string Value { get; set; } } static class MyProgram { [STAThread] static void Main(string[] args) { string xmlstr =@" <Client> <Plugin Type=""FilePlugin""> <Message>i am a file plugin</Message> <Path>c:\</Path> </Plugin> <Plugin Type=""RegsitryPlugin""> <Message>i am a registry plugin</Message> <Key>HKLM\Software\Microsoft</Key> <Name>Version</Name> <Value>3.5</Value> </Plugin> </Client> "; Assembly asm = Assembly.GetExecutingAssembly(); XDocument xDoc = XDocument.Load(new StringReader(xmlstr)); Plugin[] plugins = xDoc.Descendants("Plugin") .Select(plugin => { string typeName = plugin.Attribute("Type").Value; var type = asm.GetTypes().Where(t => t.Name == typeName).First(); Plugin p = Activator.CreateInstance(type) as Plugin; p.Type = typeName; foreach (var prop in plugin.Descendants()) { var pi = type.GetProperty(prop.Name.LocalName); object newVal = Convert.ChangeType(prop.Value, pi.PropertyType); pi.SetValue(p, newVal, null); } return p; }).ToArray(); // //"plugins" ready to use // } } } I am trying to modify the code and adding new class named DetailedPlugin so it will be able to read the following xml: string newXml = @" <Client> <Plugin Type=""FilePlugin""> <Message>i am a file plugin</Message> <Path>c:\</Path> <DetailedPlugin Type=""DetailedFilePlugin""> <Message>I am a detailed file plugin</Message> </DetailedPlugin> </Plugin> <Plugin Type=""RegsitryPlugin""> <Message>i am a registry plugin</Message> <Key>HKLM\Software\Microsoft</Key> <Name>Version</Name> <Value>3.5</Value> <DetailedPlugin Type=""DetailedRegsitryPlugin""> <Message>I am a detailed registry plugin</Message> </DetailedPlugin> </Plugin> </Client> "; for this i modified my classes to the following: public abstract class Plugin { public string Type { get; set; } public string Message { get; set; } public DetailedPlugin DetailedPlugin { get; set; } // new } public class FilePlugin : Plugin { public string Path { get; set; } } public class RegsitryPlugin : Plugin { public string Key { get; set; } public string Name { get; set; } public string Value { get; set; } } // new classes: public abstract class DetailedPlugin { public string Type { get; set; } public string Message { get; set; } } public class DetailedFilePlugin : Plugin { public string ExtraField1 { get; set; } } public class DetailedRegsitryPlugin : Plugin { public string ExtraField2{ get; set; } } from here i need some help to accomplish reading the xml and create the plugins with the nested DetailedPlugin

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  • jquery for each function

    - by jonathan p
    I have added a show more or less function to a div - this all works fine however now i need to only allow this functionality if a element is over a certain height in this case. there are numerous classes of the same so i need to do the check on each element. although i am having problems getting it to work see code below : $(document).ready(function() { $(".less").hide(); $(".more").each(function() { var actualHeight = ($(this).parent().parent().find('.appsList').height()); if (actualHeight < 150) { $(".more").hide(); } }); $(".more").click(function() { var paragraphHeight = ($(this).parent().parent().find('.appsList').height()); if (paragraphHeight > 150) { $(this).parent().parent().find('.appsHolderBody').animate({height:(paragraphHeight + 100) }); $(this).hide('slow'); $(this).parent().find('.less').show(); } return false; }); $(".less").click(function() { $(this).parent().parent().find('.appsHolderBody').animate({height:190 }); $(this).hide('slow'); $(this).parent().find('.more').show(); return false; }); }); Any help would be greatly appreciated - please note when i am targeting the parent using .parent.parent i know its not pretty but could'nt get it to run using eq(4) for some reason. so the main problem is with this part of code $(".more").each(function() { var actualHeight = ($(this).parent().parent().find('.appsList').height()); if (actualHeight < 150) { $(".more").hide(); } it hides all of the elements $(".more") instead of the ones that match the condition. html as requested <div class="appsHolder"> <div class="appsHolderBody"> <h5 class="appTitle">General Apps</h5> <ul class="appsList"> <li class="mainAppList">Resource Manager</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Resource Manager</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> <li class="mainAppList">Yet Another App</li> <li><a href="">Launch</a> <a href="">Info</a></li> </ul> </div> <div class="appsHolderExpander"> <a href="" class="more">More <img src="/wps/PA_applicationsintros/./img/downArrow.png" /></a> <a href="" class="less">Less <img src="/wps/PA_applicationsintros/./img/upArrow.png" /></a> </div> </div> <div class="appsHolderAdvertising"> <div class="appsHolderBody"> <h5 class="appTitle">Advertising Apps</h5> <ul class="appsList"> <li class="mainAppList">ATEX</li> <li><a href="">Launch</a> <a href="">Info</a></li> </ul> </div> <div class="appsHolderExpander"> <a href="" class="more">More <img src="/wps/PA_applicationsintros/./img/downArrow.png" /></a> <a href="" class="less">Less <img src="/wps/PA_applicationsintros/./img/upArrow.png" /></a> </div> </div> cheers in advance

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  • CSS Position Help (horizontal sidebar showing up when animate content over)

    - by jstacks
    Let me try my best to explain what I'd like to have happen, show you the code I have an hopefully I can get some help. So, I'm trying to do a sliding navigation UI from the left side of the screen (like a lot of mobile apps). The main content slides over, displaying the navigation menu beneath. Right now the browser thinks the screen is getting wider and introduces a horizontal scroll bar. However, I don't want that to happen... How do I get the div to animate off screen but not enlarge the width of the screen (i.e. keep it partially off screen)? Anyway here is my fiddle: http://jsfiddle.net/2vP67/6/ And here is the code within the post: HTML <div id='wrapper'> <div id='navWide'> </div> <div id='containerWide'> </div> <div id='containerTall'> <div id='container'> <div id='nav'> <div id='navNavigate'> Open Menu </div> <div id='navNavigateHide'> Close Menu </div> </div> </div> </div> <div id='sideContainerTall'> <div id='sideContainer'> <div id='sideNav'>Side Navigation </div> </div> </div> </div> CSS #wrapper { width:100%; min-width:1000px; height:100%; min-height:100%; position:relative; top:0; left:0; z-index:0; } #navWide { color: #ffffff; background:#222222; width:100%; min-width:1000px; height:45px; position:fixed; top:0; left:0; z-index:100; } #containerWide { width:100%; min-width:1000px; min-height:100%; position:absolute; top:45px; z-index:100; } #containerTall { color: #000000; background:#dadada; width:960px; min-height:100%; margin-left:-480px; position:absolute; top:0; left:50%; z-index:1000; } /***** main container *****/ #container { width:960px; min-height:585px; } #nav { color: #ffffff; background:#222222; width:960px; height:45px; position:fixed; top:0; z-index:10000; } #navNavigate { background:yellow; font-size:10px; color:#888888; width:32px; height:32px; padding:7px 6px 6px 6px; float:left; cursor:pointer; } #navNavigateHide { background:yellow; font-size:10px; color:#888888; width:32px; height:32px; padding:7px 6px 6px 6px; float:left; cursor:pointer; display:none; } #sideContainerTall { background:#888888; width:264px; min-height:100%; margin-left:-480px; position:absolute; top:0; left:50%; z-index:500; } #sideContainer { width:264px; min-height:585px; display:none; } #sideContainerTall { background:#888888; width:264px; min-height:100%; margin-left:-480px; position:absolute; top:0; left:50%; z-index:500; } #sideContainer { width:264px; min-height:585px; display:none; } #sideNav { width:264px; height:648px; float:left; } Javascript $(document).ready(function() { $('div#navNavigate').click(function() { $('div#navNavigate').hide(); $('div#navNavigateHide').show(); $('div#sideContainer').show(); $('div#containerTall').animate({ 'left': '+=264px' }); }); $('div#navNavigateHide').click(function() { $('div#navNavigate').show(); $('div#navNavigateHide').hide(); $('div#containerTall').animate({ 'left': '-=264px' }, function() { $('div#sideContainer').hide(); }); }); });

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  • strange behavior while including a class in php

    - by user1864539
    I'm experiencing a strange behavior with PHP. Basically I want to require a class within a PHP script. I know it is straight forward and I did it before but when I do so, it change the behavior of my jquery (1.8.3) ajax response. I'm running a wamp setup and my PHP version is 5.4.6. Here is a sample as for my index.html head (omitting the jquery js include) <script> $(document).ready(function(){ $('#submit').click(function(){ var action = $('#form').attr('action'); var form_data = { fname: $('#fname').val(), lname: $('#lname').val(), phone: $('#phone').val(), email: $('#email').val(), is_ajax: 1 }; $.ajax({ type: $('#form').attr('method'), url: action, data: form_data, success: function(response){ switch(response){ case 'ok': var msg = 'data saved'; break; case 'ko': var msg = 'Oops something wrong happen'; break; default: var msg = 'misc:<br/>'+response; break; } $('#message').html(msg); } }); return false; }); }); </script> body <div id="message"></div> <form id="form" action="handler.php" method="post"> <p> <input type="text" name="fname" id="fname" placeholder="fname"> <input type="text" name="lname" id="lname" placeholder="lname"> </p> <p> <input type="text" name="phone" id="phone" placeholder="phone"> <input type="text" name="email" id="email" placeholder="email"> </p> <input type="submit" name="submit" value="submit" id="submit"> </form> And as for the handler.php file: <?php require('class/Container.php'); $filename = 'xml/memory.xml'; $is_ajax = $_REQUEST['is_ajax']; if(isset($is_ajax) && $is_ajax){ $fname = $_REQUEST['fname']; $lname = $_REQUEST['lname']; $phone = $_REQUEST['phone']; $email = $_REQUEST['email']; $obj = new Container; $obj->insertData('fname',$fname); $obj->insertData('lname',$lname); $obj->insertData('phone',$phone); $obj->insertData('email',$email); $tmp = $obj->give(); $result = $tmp['_obj']; /* Push data inside array */ $array = array(); foreach($result as $key => $value){ array_push($array,$key,$value); } $xml = simplexml_load_file($filename); // check if there is any data in if(count($xml->elements->data) == 0){ // if not, create the structure $xml->elements->addChild('data',''); } // proceed now that we do have the structure if(count($xml->elements->data) == 1){ foreach($result as $key => $value){ $xml->elements->data->addChild($key,$value); } $xml->saveXML($filename); echo 'ok'; }else{ echo 'ko'; } } ? The Container class: <?php class Container{ private $_obj; public function __construct(){ $this->_obj = array(); } public function addData($data = array()){ if(!empty($data)){ $oldData = $this->_obj; $data = array_merge($oldData,$data); $this->_obj = $data; } } public function removeData($key){ if(!empty($key)){ $oldData = $this->_obj; unset($oldData[$key]); $this->_obj = $oldData; } } public function outputData(){ return $this->_obj; } public function give(){ return get_object_vars($this); } public function insertData($key,$value){ $this->_obj[$key] = $value; } } ? The strange thing is that my result always fall under the default switch statement and the ajax response fit both present statement. I noticed then if I just paste the Container class on the top of the handler.php file, everything works properly but it kind of defeat what I try to achieve. I tried different way to include the Container class but it seem to be than the issue is specific to this current scenario. I'm still learning PHP and my guess is that I'm missing something really basic. I also search on stackoverflow regarding the issue I'm experiencing as well as PHP.net, without success. Regards,

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  • Easiest way to submit data via PHP?

    - by Abijah
    I'm new to PHP, and have spent 10 hours trying to figure this problem out. The goal is to take all data entered into this order form, and send it to my email via PHP. I have 2 questions: 1. I can get PHP to send data from a single menu item (example: Mexican Tortas), but how do I get PHP to send data from multiple items (example: Mexican Tortas, Fish Sandwich and Hamburger)? 2. How do I tell PHP to not send data from menu items that don't have the "How Many?" or "Customize It?" text fields filled out? If you could provide a super simple example (or a link to a learning resource) I would really appreciate it. Thank you, Abijah PHP <?php if(isset($_POST['submit'])) { $to = "[email protected]"; $subject = "New Order"; $name_field = $_POST['name']; $phone_field = $_POST['phone']; $item = $_POST['item']; $quantity = $_POST['quantity']; $customize = $_POST['customize']; } $body = "Name: $name_field\nPhone: $phone_field\n\nItem: $item\nQuantity: $quantity\nCustomize: $customize"; echo "Data has been submitted to $to!"; mail($to, $subject, $body); ?> HTML <form action="neworder.php" method="POST"> <div class ="item"> <img style="float:left; margin-right:15px; border:1px Solid #000; width:200px; height:155px;" src="images/mexicantortas.jpg"> <h1>Mexican Torta - $8.50</h1> <input name="item" type="hidden" value="Mexican Torta"/> <h2>How Many? <font color="#999999">Ex: 1, 2, 3...?</font></h2> <input name="quantity" type="text"/> <h3>Customize It? <font color="#999999">Ex: No Lettuce, Extra Cheese...</font></h3> <textarea name="customize"/></textarea> </div><!-- ITEM_LEFT --> <div class ="item"> <img style="float:left; margin-right:15px; border:1px Solid #000; width:200px; height:155px;" src="images/fishsandwich.jpg"> <h1>Fish Sandwich - $8.50</h1> <input name="item" type="hidden" value="Fish Sandwich"/> <h2>How Many? <font color="#999999">Ex: 1, 2, 3...?</font></h2> <input name="quantity" type="text"/> <h3>Customize It? <font color="#999999">Ex: No Lettuce, Extra Cheese...</font></h3> <textarea name="customize"/></textarea> </div><!-- ITEM_LEFT --> <div class ="item"> <img style="float:left; margin-right:15px; border:1px Solid #000; width:200px; height:155px;" src="images/hamburgers.jpg"> <h1>Hamburger w/ Fries - $7.00</h1> <input name="item" type="hidden" value="Fish Sandwich"/> <h2>How Many? <font color="#999999">Ex: 1, 2, 3...?</font></h2> <input name="quantity" type="text"/> <h3>Customize It? <font color="#999999">Ex: No Lettuce, Extra Cheese...</font></h3> <textarea name="customize"/></textarea> </div><!-- ITEM_LEFT --> <div class="horizontal_form"> <div class="form"> <h2>Place Your Order Now: <font size="3"><font color="#037B41">Fill in the form below, and we'll call you when your food is ready to be picked up...</font></font></h2> <p class="name"> <input type="text" name="name" id="name" style="text-align:center;" onClick="this.value='';" value="Enter your name"/> </p> <p class="phone"> <input type="text" name="phone" id="phone" style="text-align:center;" onClick="this.value='';" value="Enter your phone #"/> </p> <p class="submit"> <input type="submit" value="Place Order" name="submit"/> </p> </div><!-- FORM --> </div><!-- HORIZONTAL_FORM --> </form>

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  • Android application displays black screen after running

    - by frgnvola
    When I click "Run as an Android Application" on Eclipse, the following is displayed in the console [2014-06-05 20:07:18 - StudentConnect] Android Launch! [2014-06-05 20:07:18 - StudentConnect] adb is running normally. [2014-06-05 20:07:18 - StudentConnect] Performing sandhu.student.connect.SplashActivity activity launch [2014-06-05 20:07:18 - StudentConnect] Using default Build Tools revision 19.0.0 [2014-06-05 20:07:18 - StudentConnect] Refreshing resource folders. [2014-06-05 20:07:18 - StudentConnect] Using default Build Tools revision 19.0.0 [2014-06-05 20:07:18 - StudentConnect] Starting incremental Pre Compiler: Checking resource changes. [2014-06-05 20:07:18 - StudentConnect] Nothing to pre compile! [2014-06-05 20:07:18 - StudentConnect] Starting incremental Package build: Checking resource changes. [2014-06-05 20:07:18 - StudentConnect] Using default Build Tools revision 19.0.0 [2014-06-05 20:07:18 - StudentConnect] Skipping over Post Compiler. [2014-06-05 20:07:20 - StudentConnect] Application already deployed. No need to reinstall. [2014-06-05 20:07:20 - StudentConnect] Starting activity sandhu.student.connect.SplashActivity on device 0f0898b2 [2014-06-05 20:07:21 - StudentConnect] ActivityManager: Starting: Intent { act=android.intent.action.MAIN cat=[android.intent.category.LAUNCHER] cmp=sandhu.student.connect/.SplashActivity } [2014-06-05 20:07:21 - StudentConnect] ActivityManager: Warning: Activity not started, its current task has been brought to the front After deployed to my phone, it only displays a black screen. I recently implemented a splash screen, but it was working fine before; however I think it might have something to do with the problem. Here are my java and xml files: MainActivity.java package sandhu.student.connect; import android.app.Activity; import android.os.Bundle; import android.view.KeyEvent; import android.view.View; import android.webkit.WebSettings; import android.webkit.WebView; import android.webkit.WebViewClient; public class MainActivity extends Activity { public WebView student_zangle; @Override protected void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.activity_main); WebView student_zangle = (WebView) findViewById(R.id.student_zangle); student_zangle.loadUrl("https://zangleweb01.clovisusd.k12.ca.us/studentconnect/"); student_zangle.setWebViewClient(new WebViewClient()); student_zangle.setScrollBarStyle(View.SCROLLBARS_INSIDE_OVERLAY); WebSettings settings = student_zangle.getSettings(); settings.setJavaScriptEnabled(true); settings.setBuiltInZoomControls(true); settings.setLoadWithOverviewMode(true); settings.setUseWideViewPort(true); } @Override public boolean onKeyDown(int keyCode, KeyEvent event) { WebView student_zangle = (WebView) findViewById(R.id.student_zangle); if ((keyCode == KeyEvent.KEYCODE_BACK) && student_zangle.canGoBack()) { student_zangle.goBack(); return true; } else { finish(); } return super.onKeyDown(keyCode, event); } } activity_main.xml <?xml version="1.0" encoding="utf-8"?> <RelativeLayout xmlns:android="http://schemas.android.com/apk/res/android" xmlns:tools="http://schemas.android.com/tools" android:layout_width="match_parent" android:layout_height="match_parent" android:background="@drawable/blue" tools:context=".MainActivity" > <WebView android:id="@+id/student_zangle" android:layout_width="match_parent" android:layout_height="match_parent" /> </RelativeLayout> SplashActivity.java package sandhu.student.connect; import android.os.Bundle; import android.preference.PreferenceActivity; public class SplashActivity extends PreferenceActivity { @SuppressWarnings("deprecation") @Override protected void onCreate(Bundle savedInstanceState) { // TODO Auto-generated method stub super.onCreate(savedInstanceState); addPreferencesFromResource(R.xml.prefs); } } splash_activity.xml <?xml version="1.0" encoding="utf-8"?> <RelativeLayout xmlns:android="http://schemas.android.com/apk/res/android" android:layout_width="match_parent" android:layout_height="match_parent" android:background="@drawable/blue" android:orientation="vertical" > <ImageView android:id="@+id/imageView1" android:layout_width="250dp" android:layout_height="100dp" android:layout_alignParentTop="true" android:layout_centerHorizontal="true" android:layout_marginTop="145dp" android:contentDescription="@string/zangle_logo" android:src="@drawable/logo" /> </RelativeLayout> Also, here is a full copy of the logcat error output: 06-05 20:19:46.698: E/Watchdog(817): !@Sync 1952 06-05 20:20:09.971: E/memtrack(16438): Couldn't load memtrack module (No such file or directory) 06-05 20:20:09.971: E/android.os.Debug(16438): failed to load memtrack module: -2 06-05 20:20:11.012: E/memtrack(16451): Couldn't load memtrack module (No such file or directory) 06-05 20:20:11.012: E/android.os.Debug(16451): failed to load memtrack module: -2 06-05 20:20:11.202: E/EnterpriseContainerManager(817): ContainerPolicy Service is not yet ready!!! Please help me figure out what is wrong, or at least point me in the right direction. Thanks in advance.

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  • $.post is not working

    - by BEBO
    i am trying to post data to Mysql using jquery $.post and php page. my code is not running and nothing is added to the mysql table. I am not sure if the path i am creating is wrong but any help would be appreciated. Jquery location: f_js/tasks/TaskTest.js <script type="text/javascript"> $(document).ready(function(){ $("#AddTask").click(function(){ var acct = $('#acct').val(); var quicktask = $('#quicktask').val(); var user = $('#user').val(); $.post('addTask.php',{acct:acct,quicktask:quicktask,user:user}, function(data){ $('#result').fadeIn('slow').html(data); }); }); }); </script> addTask.php (runs the jqeury code) <?php include 'dbconnect.php'; include 'sessions.php'; $acct = $_POST['acct']; $task = $_POST['quicktask']; $taskstatus = 'Active'; //get task Creator $user = $_POST['user']; //query task creator from users table $allusers = mysql_query("SELECT * FROM users WHERE username = '$user'"); while ($rows = mysql_fetch_array($allusers)) { //get first and last name for task creator $taskOwner = $rows['user_firstname']; $taskOwnerLast = $rows['user_lastname']; $taskOwnerFull = $taskOwner." ".$taskOwnerLast; mysql_query("INSERT INTO tasks (taskresource, tasktitle, taskdetail, taskstatus, taskowner, taskOwnerFullName) VALUES ('$acct', '$task', '$task', '$taskstatus', '$user', '$taskOwnerFull' )"); echo "inserted"; } ?> Accountview.php finally the front page <html> <div class="input-cont "> <input type="text" class="form-control col-lg-12" placeholder="Add a quick Task..." name ="quicktask" id="quicktask"> </div> <div class="form-group"> <div class="pull-right chat-features"> <a href="javascript:;"> <i class="icon-camera"></i> </a> <a href="javascript:;"> <i class="icon-link"></i> </a> <input type="button" class="btn btn-danger" name="AddTask" id="AddTask" value="Add" /> <input type="hidden" name="acct" id="acct" value="<?php echo $_REQUEST['acctname']?>"/> <input type="hidden" name="user" id="user" value="<?php $username = $_SESSION['username']; echo $username?>"/> <div id="result">result</div> </div> </div> <!-- js placed at the end of the document so the pages load faster --> <script src="js/jquery.js"></script> <script src="f_js/tasks/TaskTest.js"></script> <!--common script for all pages--> <script src="js/common-scripts.js"></script> <script type="text/javascript" src="assets/gritter/js/jquery.gritter.js"></script> <script src="js/gritter.js" type="text/javascript"></script> <script> </html> Firebug reponse: Response Headers Connection Keep-Alive Content-Length 0 Content-Type text/html Date Fri, 08 Nov 2013 21:48:50 GMT Keep-Alive timeout=5, max=100 Server Apache/2.4.4 (Win32) OpenSSL/0.9.8y PHP/5.4.16 X-Powered-By PHP/5.4.16 refresh 5; URL=index.php Request Headers Accept */* Accept-Encoding gzip, deflate Accept-Language en-US,en;q=0.5 Content-Length 13 Content-Type application/x-www-form-urlencoded; charset=UTF-8 Cookie PHPSESSID=6gufl3guiiddreg8cdlc0htnc6 Host localhost Referer http://localhost/betahtml/AccountView.php?acctname=client%201 User-Agent Mozilla/5.0 (Windows NT 6.3; WOW64; rv:25.0) Gecko/20100101 Firefox/25.0 X-Requested-With XMLHttpRequest

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  • JQuery tab Selection problem?

    - by PeAk
    New to JQuery and I was wondering how do I keep any tabbed selected when a user reloads the web page? What part of my code do I need to change? Here is my JQuery code. $(document).ready(function() { //When page loads... $(".form-content").hide(); //Hide all content var firstMenu = $("#home-menu ul li:first"); firstMenu.show(); firstMenu.find("a").addClass("selected-link"); //Activate first tab $(".form-content:first").show(); //Show first tab content //On Click Event $("#home-menu ul li").click(function() { $("#home-menu ul li a").removeClass("selected-link"); //Remove any "selected-link" class $(this).find("a").addClass("selected-link"); //Add "selected-link" class to selected tab $(".form-content").hide(); //Hide all tab content var activeTab = $(this).find("a").attr("href"); //Find the href attribute value to identify the selected-link tab + content $(activeTab).fadeIn(); //Fade in the selected-link ID content return false; }); }); Here is the XHTML code. <div id="home-menu"> <ul> <li><a href="#personal-info-form" title="Personal Info Form Link">Personal Info</a></li> <li><a href="#contact-info-form" title="Contact Info Form Link">Contact Info</a></li> </ul> </div> <div> <div id="personal-info-form" class="form-content"> <h2>Personal Information</h2> <form method="post" action="index.php"> <fieldset> <ul> <li><label for="first_name">First Name: </label><input type="text" name="first_name" id="first_name" size="25" class="input-size" value="<?php if(!empty($first_name)){ echo $first_name; } ?>" /></li> <li><label for="middle_name">Middle Name: </label><input type="text" name="middle_name" id="middle_name" size="25" class="input-size" value="<?php if(!empty($middle_name)){ echo $middle_name; } ?>" /></li> <li><label for="last_name">Last Name: </label><input type="text" name="last_name" id="last_name" size="25" class="input-size" value="<?php if(!empty($last_name)){ echo $last_name; } ?>" /></li> <li><label for="password-1">Password: </label><input type="password" name="password1" id="password-1" size="25" class="input-size" /></li> <li><label for="password-2">Confirm Password: </label><input type="password" name="password2" id="password-2" size="25" class="input-size" /></li> <li><input type="submit" name="submit" value="Save Changes" class="save-button" /> <input type="submit" name="submit" value="Preview Changes" class="preview-changes-button" /></li> </ul> </fieldset> </form> </div> <div id="contact-info-form" class="form-content"> <h2>Contact Information</h2> <form method="post" action="index.php" id="contact-form"> <fieldset> <ul> <li><label for="address">Address 1: </label><input type="text" name="address" id="address" size="25" class="input-size" value="<?php if (isset($_POST['address'])) { echo $_POST['address']; } else if(!empty($address)) { echo $address; } ?>" /></li> <li><label for="address_two">Address 2: </label><input type="text" name="address_two" id="address_two" size="25" class="input-size" value="<?php if (isset($_POST['address_two'])) { echo $_POST['address_two']; } else if(!empty($address_two)) { echo $address_two; } ?>" /></li> <li><label for="city_town">City/Town: </label><input type="text" name="city_town" id="city_town" size="25" class="input-size" value="<?php if (isset($_POST['city_town'])) { echo $_POST['city_town']; } else if(!empty($city_town)) { echo $city_town; } ?>" /></li> <li><input type="submit" name="submit" value="Save Changes" class="save-button" /> <input type="hidden" name="contact_info_submitted" value="true" /> <input type="submit" name="submit" value="Preview Changes" class="preview-changes-button" /></li> </ul> </fieldset> </form> </div> </div>

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  • How to specify 2 different positions for Colorbox or Fancybox on the same page?

    - by Eric
    I know this question has been asked before, but I'm having difficulty implementing it. I'm looking for a more specific answer. Here is my html code: <head> <meta charset=utf-8 /> <title>ColorBox Examples</title> <style type="text/css"> body{font:12px/1.2 Verdana, Arial, san-serrif; padding:0 10px;} a:link, a:visited{text-decoration:none; color:#416CE5; border-bottom:1px solid #416CE5;} h2{font-size:13px; margin:15px 0 0 0;} </style> <link media="screen" rel="stylesheet" href="colorbox.css" /> <script src="http://ajax.googleapis.com/ajax/libs/jquery/1.4.4/jquery.min.js"></script> <script src="../colorbox/jquery.colorbox.js"></script> <script> $(document).ready(function(){ //Examples of how to assign the ColorBox event to elements $(".example7").colorbox({width:"80%", height:"80%", iframe:true}); $(".example7").colorbox({width:"80%", height:"80%", iframe:true}); }); </script> </head> <body> <p><a class='example7' href="http://google.com">Outside Webpage 1 (Iframe)</a></p> <p><a class='example7' href="http://google.com">Outside Webpage 2 (Iframe)</a></p> (Excuse the wrong indentation - I had to mess with the formatting to get the body content to show up.) Here is my CSS code(default colorbox code): #colorbox, #cboxOverlay, #cboxWrapper{position:absolute; top:0; left:0; z-index:9999; overflow:hidden;} #cboxOverlay{position:fixed; width:100%; height:100%;} #cboxMiddleLeft, #cboxBottomLeft{clear:left;} #cboxContent{position:relative; overflow:visible;} #cboxLoadedContent{overflow:auto;} #cboxLoadedContent iframe{display:block; width:100%; height:100%; border:0;} #cboxTitle{margin:0;} #cboxLoadingOverlay, #cboxLoadingGraphic{position:absolute; top:0; left:0; width:100%;} #cboxPrevious, #cboxNext, #cboxClose, #cboxSlideshow{cursor:pointer;} #cboxOverlay{background:#fff;} #colorbox{} #cboxContent{margin-top:32px;} #cboxLoadedContent{background:#000; padding:1px;} #cboxLoadingGraphic{background:url(images/loading.gif) no-repeat center center;} #cboxLoadingOverlay{background:#000;} #cboxTitle{position:absolute; top:-22px; left:0; color:#000;} #cboxCurrent{position:absolute; top:-22px; right:205px; text-indent:-9999px;} #cboxSlideshow, #cboxPrevious, #cboxNext, #cboxClose{text-indent:-9999px; width:20px; height:20px; position:absolute; top:-20px; background:url(images/controls.png) no-repeat 0 0;} #cboxPrevious{background-position:0px 0px; right:44px;} #cboxPrevious.hover{background-position:0px -25px;} #cboxNext{background-position:-25px 0px; right:22px;} #cboxNext.hover{background-position:-25px -25px;} #cboxClose{background-position:-50px 0px; right:0;} #cboxClose.hover{background-position:-50px -25px;} .cboxSlideshow_on #cboxPrevious, .cboxSlideshow_off #cboxPrevious{right:66px;} .cboxSlideshow_on #cboxSlideshow{background-position:-75px -25px; right:44px;} .cboxSlideshow_on #cboxSlideshow.hover{background-position:-100px -25px;} .cboxSlideshow_off #cboxSlideshow{background-position:-100px 0px; right:44px;} .cboxSlideshow_off #cboxSlideshow.hover{background-position:-75px -25px;} Can someone please tell me how this can be achieved? Forgive my lack of css knowledge :) Any help would be greatly appreciated. Thanks a ton.

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  • Building my first Javascript Application (jQuery), struggling on something

    - by Jason Wells
    I'd really appreciate recommendations on the most efficient way to approach this. I'm building a simple javascript application which displays a list of records and allows the user to edit a record by clicking an "Edit" link in the records row. The user also can click the "Add" link to pop open a dialog allowing them to add a new record. Here's a working prototype of this: http://jsfiddle.net/FfRcG/ You'll note if you click "Edit" a dialog pops up with some canned values. And, if you click "Add", a dialog pops up with empty values. I need help on how to approach two problems I believe we need to pass our index to our edit dialog and reference the values within the JSON, but I am unsure how to pass the index when the user clicks edit. It bothers me that the Edit and Add div contents are so similiar (Edit just pre populates the values). I feel like there is a more efficient way of doing this but am at a loss. Here is my code for reference $(document).ready( function(){ // Our JSON (This would actually be coming from an AJAX database call) people = { "COLUMNS":["DATEMODIFIED", "NAME","AGE"], "DATA":[ ["9/6/2012", "Person 1","32"], ["9/5/2012","Person 2","23"] ] } // Here we loop over our JSON and build our HTML (Will refactor to use templating eventually) members = people.DATA; var newcontent = '<table width=50%><tr><td>date</td><td>name</td><td>age</td><td></td></tr>'; for(var i=0;i<members.length;i++) { newcontent+= '<tr id="member'+i+'"><td>' + members[i][0] + '</td>'; newcontent+= '<td>' + members[i][1] + '</td>'; newcontent+= '<td>' + members[i][2] + '</td>'; newcontent+= '<td><a href="#" class="edit" id=edit'+i+'>Edit</a></td><td>'; } newcontent += "</table>"; $("#result").html(newcontent); // Bind a dialog to the edit link $(".edit").click( function(){ // Trigger our dialog to open $("#edit").dialog("open"); // Not sure the most efficient way to change our dialog field values $("#name").val() // ??? alert($()); return false; }); // Bind a dialog to the add link $(".edit").click( function(){ // Trigger our dialog to open $("#add").dialog("open"); return false; }); // Bind a dialog to our edit DIV $("#edit").dialog(); // Bind a dialog to our add DIV $("#add").dialog(); }); And here's the HTML <h1>People</h1> <a href="#" class="add">Add a new person</a> <!-- Where results show up --> <div id="result"></div> <!-- Here's our edit DIV - I am not clear as to the best way to pass the index in our JSON so that we can reference positions in our array to pre populate the input values. --> <div id="edit"> <form> <p>Name:<br/><input type="text" id="name" value="foo"></p> <p>Age:<br/><input type="text" id="age" value="33"></p> <input type="submit" value="Save" id="submitEdit"> </form> </div> <!-- Here's our add DIV - This layout is so similiar to our edit dialog. What is the most efficient way to handle a situation like this? --> <div id="add"> <form> <p>Name:<br/><input type="text" id="name"></p> <p>Age:<br/><input type="text" id="age"></p> <input type="submit" value="Save" id="submitEdit"> </form> </div>

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  • Cannot Start Nginx Compiled from Source

    - by Jason Alan Kennedy
    I am trying to compile Nginx from source based on the original compiled Nginx server running on my DigitalOcean server ( Ubuntu-14.04 64x ) but with a few extra modules. I can get everything installed smoothly but I can not get it to start. I am sure the ini is correct because I copied the original source off the current running Nginx server [ Even though I see that Nginx now adds the ini when compiling fron source ]. Below is the [ lengthy process ] that I am performing - add sorry but I wanted to be thorough for those who are in need of the info ]. Because I am a newB to Nginx, I am sure I am missing something or just have it all wrong. If you may look over what I have done and see if you spot anything I need/need to change, I will greatly appreciate it. Thnx! With the original Nginx server still running: I check the current/running Nginx configuration so I can build the new Nginx instance the same but with the added modules: nginx -V # The out-put: configure arguments: --with-cc-opt='-g -O2 -fstack-protector --param=ssp-buffer-size=4 -Wformat -Werror=format-security -D_FORTIFY_SOURCE=2' --with-ld-opt='-Wl,-Bsymbolic-functions -Wl,-z,relro' --prefix=/usr/share/nginx --conf-path=/etc/nginx/nginx.conf --http-log-path=/var/log/nginx/access.log --error-log-path=/var/log/nginx/error.log --lock-path=/var/lock/nginx.lock --pid-path=/run/nginx.pid --http-client-body-temp-path=/var/lib/nginx/body --http-fastcgi-temp-path=/var/lib/nginx/fastcgi --http-proxy-temp-path=/var/lib/nginx/proxy --http-scgi-temp-path=/var/lib/nginx/scgi --http-uwsgi-temp-path=/var/lib/nginx/uwsgi --with-debug --with-pcre-jit --with-ipv6 --with-http_ssl_module --with-http_stub_status_module --with-http_realip_module --with-http_addition_module --with-http_dav_module --with-http_geoip_module --with-http_gzip_static_module --with-http_image_filter_module --with-http_spdy_module --with-http_sub_module --with-http_xslt_module NOTE: The configure arguments below return errors during 'make' so I removed them. I don't know what they are - could this be related to my issue??? --with-cc-opt='-g -O2 -fstack-protector --param=ssp-buffer-size=4 -Wformat -Werror=format-security -D_FORTIFY_SOURCE=2' --with-ld-opt='-Wl,-Bsymbolic-functions -Wl,-z,relro' Moving on: # So I don't have to sudo every line: sudo bash # Check for updates first thing: apt-get update # Install various prerequisites needed to compile Nginx: apt-get install build-essential libgd2-xpm-dev lsb-base zlib1g-dev libpcre3 libpcre3-dev libbz2-dev libxslt1-dev libxml2 libssl-dev libgeoip-dev tar unzip openssl # Create System users [ if it doesn't exist - but I see its there on DigitalOceans' Droplets all-ready ]: adduser --system --no-create-home --disabled-login --disabled-password --group www-data # Download NGINX wget http://nginx.org/download/nginx-1.7.4.tar.gz tar -xvzf nginx-1.7.4.tar.gz # Then Google PageSpeed: wget https://github.com/pagespeed/ngx_pagespeed/archive/release-1.8.31.4-beta.zip unzip release-1.8.31.4-beta.zip # cd into the PageSpeed Directory cd ngx_pagespeed-release-1.8.31.4-beta/ # and add the PSOL files in there: wget https://dl.google.com/dl/page-speed/psol/1.8.31.4.tar.gz tar -xzvf 1.8.31.4.tar.gz # Get back to the root directory: cd # I add the ngx_cache_purge module and will install the Nginx Helper plugin for WP later: wget https://github.com/FRiCKLE/ngx_cache_purge/archive/2.1.zip unzip 2.1.zip # Add the headers-more-nginx-module: wget https://github.com/openresty/headers-more-nginx-module/archive/v0.25.zip unzip v0.25.zip # and the naxsi module for added security: wget https://github.com/nbs-system/naxsi/archive/0.53-2.tar.gz tar -xvzf 0.53-2.tar.gz # cd to the new Nginx directory cd nginx-1.7.4 # Set up the configuration build based on the current running Nginx config args and add my additional modules: ./configure \ --add-module=$HOME/naxsi-0.53-2/naxsi_src \ --prefix=/usr/share/nginx \ --conf-path=/etc/nginx/nginx.conf \ --http-log-path=/var/log/nginx/access.log \ --error-log-path=/var/log/nginx/error.log \ --lock-path=/var/lock/nginx.lock \ --pid-path=/run/nginx.pid \ --http-client-body-temp-path=/var/lib/nginx/body \ --http-fastcgi-temp-path=/var/lib/nginx/fastcgi \ --http-proxy-temp-path=/var/lib/nginx/proxy \ --http-scgi-temp-path=/var/lib/nginx/scgi \ --http-uwsgi-temp-path=/var/lib/nginx/uwsgi \ --user=www-data \ --group=www-data \ --with-debug \ --with-pcre-jit \ --with-ipv6 \ --with-http_ssl_module \ --with-http_stub_status_module \ --with-http_realip_module \ --with-http_addition_module \ --with-http_dav_module \ --with-http_geoip_module \ --with-http_gzip_static_module \ --with-http_image_filter_module \ --with-http_spdy_module \ --with-http_sub_module \ --with-http_xslt_module \ --with-mail \ --with-mail_ssl_module \ --add-module=$HOME/ngx_pagespeed-release-1.8.31.4-beta \ --add-module=$HOME/ngx_cache_purge-2.1 \ --add-module=$HOME/headers-more-nginx-module-0.25 [ENTER] Configuration Summary: Configuration summary + using system PCRE library + using system OpenSSL library + md5: using OpenSSL library + sha1: using OpenSSL library + using system zlib library nginx path prefix: "/usr/share/nginx" nginx binary file: "/usr/share/nginx/sbin/nginx" nginx configuration prefix: "/etc/nginx" nginx configuration file: "/etc/nginx/nginx.conf" nginx pid file: "/run/nginx.pid" nginx error log file: "/var/log/nginx/error.log" nginx http access log file: "/var/log/nginx/access.log" nginx http client request body temporary files: "/var/lib/nginx/body" nginx http proxy temporary files: "/var/lib/nginx/proxy" nginx http fastcgi temporary files: "/var/lib/nginx/fastcgi" nginx http uwsgi temporary files: "/var/lib/nginx/uwsgi" nginx http scgi temporary files: "/var/lib/nginx/scgi" Next step: I cd to root and I check the old Nginx folder locations and double checked the 'make' output to see that they are the same: whereis nginx #Output: nginx: /usr/sbin/nginx /etc/nginx /usr/share/nginx NOTE: Not sure about the '/usr/sbin/nginx' - Possible issue??? Next I copy the old /etc/nginx/nginx.conf, /etc/nginx/sites-available/default, /etc/nginx/sites-enabled/default, /etc/init.d/nginx to a text file locally for safe keeping to use in the new Nginx server. Then stop the running Nginx server: service nginx stop , verify it's stopped: service --status-all and the output is: [ - ] nginx To verify that there are two Nginx directories, I cd to: cd nginx* and the output is an error indicating there are two nginx folders - Cool Beans! :) Now Install the new Nginx server: cd nginx-1.7.4 make install # INSTALL OUTPUT ######################################## make -f objs/Makefile install make[1]: Entering directory `/home/walkingfish/nginx-1.7.4' test -d '/usr/share/nginx' || mkdir -p '/usr/share/nginx' test -d '/usr/share/nginx/sbin' || mkdir -p '/usr/share/nginx/sbin' test ! -f '/usr/share/nginx/sbin/nginx' || mv '/usr/share/nginx/sbin/nginx' '/usr/share/nginx/sbin/nginx.old' cp objs/nginx '/usr/share/nginx/sbin/nginx' test -d '/etc/nginx' || mkdir -p '/etc/nginx' cp conf/koi-win '/etc/nginx' cp conf/koi-utf '/etc/nginx' cp conf/win-utf '/etc/nginx' test -f '/etc/nginx/mime.types' || cp conf/mime.types '/etc/nginx' cp conf/mime.types '/etc/nginx/mime.types.default' test -f '/etc/nginx/fastcgi_params' || cp conf/fastcgi_params '/etc/nginx' cp conf/fastcgi_params '/etc/nginx/fastcgi_params.default' test -f '/etc/nginx/fastcgi.conf' || cp conf/fastcgi.conf '/etc/nginx' cp conf/fastcgi.conf '/etc/nginx/fastcgi.conf.default' test -f '/etc/nginx/uwsgi_params' || cp conf/uwsgi_params '/etc/nginx' cp conf/uwsgi_params '/etc/nginx/uwsgi_params.default' test -f '/etc/nginx/scgi_params' || cp conf/scgi_params '/etc/nginx' cp conf/scgi_params '/etc/nginx/scgi_params.default' test -f '/etc/nginx/nginx.conf' || cp conf/nginx.conf '/etc/nginx/nginx.conf' cp conf/nginx.conf '/etc/nginx/nginx.conf.default' test -d '/run' || mkdir -p '/run' test -d '/var/log/nginx' || mkdir -p '/var/log/nginx' test -d '/usr/share/nginx/html' || cp -R html '/usr/share/nginx' test -d '/var/log/nginx' || mkdir -p '/var/log/nginx' ######################################################### I copy/create the files that I saved earlier to txt files in sites-available, the config, default and ini files then symlink them to sites-enabled, and so on. And now to start the server: service nginx start And this is where s#!+ hits the fan - Nada. I check to see if Nginx is running with service --status-all and its not. Also with nginx -V and its not installed??? I reboot the system too and still nothing. So I am not sure what is wrong here. The ini was copied over from the old server along with all the other config files after deleting the old files. When I opened the new compiled files, the nginx default data was present so I replaced them with my old original data prior to starting the new server for the first time. Also to be safe, I rm /etc/nginx/sites-enabled/default and symlinked with ln -s /etc/nginx/sites-available/default /etc/nginx/sites-enabled/default with no errors and I verified that the data was in the sites-enabled/default file. I don't think the server really/fully installed because of the nginx -V result: The program 'nginx' can be found in the following packages: * nginx-core * nginx-extras * nginx-full * nginx-light * nginx-naxsi Try: apt-get install <selected package> Do/should I apt-get install nginx-1.7.4 ?? Or what package do I use being that its a custom package and make install earlier did nothing?? If you need to see the conf files I copied over from the old to the custom server, LMK and I'll post them. Again your help here would be appreciated!

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  • Mysql server crashes Innodb

    - by martin
    Today we got some DB crash. The DB is InnoDB. At firstin log: 120404 10:57:40 InnoDB: ERROR: the age of the last checkpoint is 9433732, InnoDB: which exceeds the log group capacity 9433498. InnoDB: If you are using big BLOB or TEXT rows, you must set the InnoDB: combined size of log files at least 10 times bigger than the InnoDB: largest such row. 120404 10:58:48 InnoDB: ERROR: the age of the last checkpoint is 9825579, InnoDB: which exceeds the log group capacity 9433498. InnoDB: If you are using big BLOB or TEXT rows, you must set the InnoDB: combined size of log files at least 10 times bigger than the InnoDB: largest such row. 120404 10:59:04 InnoDB: ERROR: the age of the last checkpoint is 13992586, InnoDB: which exceeds the log group capacity 9433498. InnoDB: If you are using big BLOB or TEXT rows, you must set the InnoDB: combined size of log files at least 10 times bigger than the InnoDB: largest such row. 120404 10:59:20 InnoDB: ERROR: the age of the last checkpoint is 18059881, InnoDB: which exceeds the log group capacity 9433498. InnoDB: If you are using big BLOB or TEXT rows, you must set the InnoDB: combined size of log files at least 10 times bigger than the InnoDB: largest such row. after manual service stop and normal PC restart : 120404 11:12:35 InnoDB: Error: page 3473451 log sequence number 105 802365904 InnoDB: is in the future! Current system log sequence number 105 796344770. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. InnoDB: 1 transaction(s) which must be rolled back or cleaned up InnoDB: in total 1 row operations to undo InnoDB: Trx id counter is 0 1103869440 120404 11:12:37 InnoDB: Error: page 0 log sequence number 105 834817616 InnoDB: is in the future! Current system log sequence number 105 796344770. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. InnoDB: Last MySQL binlog file position 0 3710603, file name .\mysql-bin.000336 InnoDB: Starting in background the rollback of uncommitted transactions 120404 11:12:38 InnoDB: Rolling back trx with id 0 1103866646, 1 rows to undo 120404 11:12:38 InnoDB: Started; log sequence number 105 796344770 120404 11:12:38 InnoDB: Error: page 2097163 log sequence number 105 803249754 InnoDB: is in the future! Current system log sequence number 105 796344770. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. InnoDB: Rolling back of trx id 0 1103866646 completed 120404 11:12:39 InnoDB: Rollback of non-prepared transactions completed 120404 11:12:39 [Note] Event Scheduler: Loaded 0 events 120404 11:12:39 [Note] wampmysqld: ready for connections. Version: '5.1.53-community' socket: '' port: 3306 MySQL Community Server (GPL) 120404 11:12:40 InnoDB: Error: page 2097162 log sequence number 105 803215859 InnoDB: is in the future! Current system log sequence number 105 796345097. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. 120404 11:12:40 InnoDB: Error: page 2097156 log sequence number 105 803181181 InnoDB: is in the future! Current system log sequence number 105 796345097. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. 120404 11:12:40 InnoDB: Error: page 2097157 log sequence number 105 803193066 InnoDB: is in the future! Current system log sequence number 105 796345097. InnoDB: Your database may be corrupt or you may have copied the InnoDB InnoDB: tablespace but not the InnoDB log files. See InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: for more information. when tried to recover data get : key_buffer_size=16777216 read_buffer_size=262144 max_used_connections=0 max_threads=151 threads_connected=0 It is possible that mysqld could use up to key_buffer_size + (read_buffer_size + sort_buffer_size)*max_threads = 133725 K bytes of memory Hope that's ok; if not, decrease some variables in the equation. thd: 0x0 Attempting backtrace. You can use the following information to find out where mysqld died. If you see no messages after this, something went terribly wrong... 0000000140262AFC mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402AAFA1 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402AB33A mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 0000000140268219 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014027DB13 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402A909F mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402A91B6 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014025B9B0 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014022F9C6 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 0000000140219979 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014009ABCF mysqld.exe!?ha_initialize_handlerton@@YAHPEAUst_plugin_int@@@Z() 000000014003308C mysqld.exe!?plugin_lock_by_name@@YAPEAUst_plugin_int@@PEAVTHD@@PEBUst_mysql_lex_string@@H@Z() 00000001400375A9 mysqld.exe!?plugin_init@@YAHPEAHPEAPEADH@Z() 000000014001DACE mysqld.exe!handle_shutdown() 000000014001E285 mysqld.exe!?win_main@@YAHHPEAPEAD@Z() 000000014001E632 mysqld.exe!?mysql_service@@YAHPEAX@Z() 00000001402EA477 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402EA545 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000007712652D kernel32.dll!BaseThreadInitThunk() 000000007725C521 ntdll.dll!RtlUserThreadStart() The manual page at http://dev.mysql.com/doc/mysql/en/crashing.html contains information that should help you find out what is causing the crash. 120404 14:17:49 [Note] Plugin 'FEDERATED' is disabled. 120404 14:17:49 [Warning] option 'innodb-force-recovery': signed value 8 adjusted to 6 InnoDB: The user has set SRV_FORCE_NO_LOG_REDO on InnoDB: Skipping log redo InnoDB: Error: trying to access page number 4290979199 in space 0, InnoDB: space name .\ibdata1, InnoDB: which is outside the tablespace bounds. InnoDB: Byte offset 0, len 16384, i/o type 10. InnoDB: If you get this error at mysqld startup, please check that InnoDB: your my.cnf matches the ibdata files that you have in the InnoDB: MySQL server. 120404 14:17:52 InnoDB: Assertion failure in thread 3928 in file .\fil\fil0fil.c lin23 InnoDB: We intentionally generate a memory trap. InnoDB: Submit a detailed bug report to http://bugs.mysql.com. InnoDB: If you get repeated assertion failures or crashes, even InnoDB: immediately after the mysqld startup, there may be InnoDB: corruption in the InnoDB tablespace. Please refer to InnoDB: http://dev.mysql.com/doc/refman/5.1/en/forcing-recovery.html InnoDB: about forcing recovery. 120404 14:17:52 - mysqld got exception 0xc0000005 ; This could be because you hit a bug. It is also possible that this binary or one of the libraries it was linked against is corrupt, improperly built, or misconfigured. This error can also be caused by malfunctioning hardware. We will try our best to scrape up some info that will hopefully help diagnose the problem, but since we have already crashed, something is definitely wrong and this may fail. key_buffer_size=16777216 read_buffer_size=262144 max_used_connections=0 max_threads=151 threads_connected=0 It is possible that mysqld could use up to key_buffer_size + (read_buffer_size + sort_buffer_size)*max_threads = 133725 K bytes of memory Hope that's ok; if not, decrease some variables in the equation. thd: 0x0 Attempting backtrace. You can use the following information to find out where mysqld died. If you see no messages after this, something went terribly wrong... 0000000140262AFC mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402AAFA1 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402AB33A mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 0000000140268219 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014027DB13 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402A909F mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402A91B6 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014025B9B0 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014022F9C6 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 0000000140219979 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000014009ABCF mysqld.exe!?ha_initialize_handlerton@@YAHPEAUst_plugin_int@@@Z() 000000014003308C mysqld.exe!?plugin_lock_by_name@@YAPEAUst_plugin_int@@PEAVTHD@@PEBUst_mysql_lex_string@@H@Z() 00000001400375A9 mysqld.exe!?plugin_init@@YAHPEAHPEAPEADH@Z() 000000014001DACE mysqld.exe!handle_shutdown() 000000014001E285 mysqld.exe!?win_main@@YAHHPEAPEAD@Z() 000000014001E632 mysqld.exe!?mysql_service@@YAHPEAX@Z() 00000001402EA477 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 00000001402EA545 mysqld.exe!?check_next_symbol@Gis_read_stream@@QEAA_ND@Z() 000000007712652D kernel32.dll!BaseThreadInitThunk() 000000007725C521 ntdll.dll!RtlUserThreadStart() The manual page at http://dev.mysql.com/doc/mysql/en/crashing.html contains information that should help you find out what is causing the crash. Any suggestion how to get DB working ????

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  • socket connection failed, telnet OK

    - by cf16
    my problem is that I can't connect two comps through socket (windows xp and windows7) although the server created with socket is listening and I can telnet it. It receives then information and does what should be done, but if I run the corresponding socket client I get error 10061. Moreover I am behind firewall - these two comps are running within my LAN, the windows firewalls are turned off, comp1: 192.168.1.2 port 12345 comp1: 192.168.1.6 port 12345 router: 192.168.1.1 Maybe port forwarding could help? But most important for me is to answer why Sockets fail if telnet works fine. client: int main(){ // Initialize Winsock. WSADATA wsaData; int iResult = WSAStartup(MAKEWORD(2,2), &wsaData); if (iResult != NO_ERROR) printf("Client: Error at WSAStartup().\n"); else printf("Client: WSAStartup() is OK.\n"); // Create a socket. SOCKET m_socket; m_socket = socket(AF_INET, SOCK_STREAM, IPPROTO_TCP); if (m_socket == INVALID_SOCKET){ printf("Client: socket() - Error at socket(): %ld\n", WSAGetLastError()); WSACleanup(); return 7; }else printf("Client: socket() is OK.\n"); // Connect to a server. sockaddr_in clientService; clientService.sin_family = AF_INET; //clientService.sin_addr.s_addr = inet_addr("77.64.240.156"); clientService.sin_addr.s_addr = inet_addr("192.168.1.5"); //clientService.sin_addr.s_addr = inet_addr("87.207.222.5"); clientService.sin_port = htons(12345); if (connect(m_socket, (SOCKADDR*)&clientService, sizeof(clientService)) == SOCKET_ERROR){ printf("Client: connect() - Failed to connect.\n"); wprintf(L"connect function failed with error: %ld\n", WSAGetLastError()); iResult = closesocket(m_socket); if (iResult == SOCKET_ERROR) wprintf(L"closesocket function failed with error: %ld\n", WSAGetLastError()); WSACleanup(); return 6; } // Send and receive data int bytesSent; int bytesRecv = SOCKET_ERROR; // Be careful with the array bound, provide some checking mechanism char sendbuf[200] = "Client: Sending some test string to server..."; char recvbuf[200] = ""; bytesSent = send(m_socket, sendbuf, strlen(sendbuf), 0); printf("Client: send() - Bytes Sent: %ld\n", bytesSent); while(bytesRecv == SOCKET_ERROR){ bytesRecv = recv(m_socket, recvbuf, 32, 0); if (bytesRecv == 0 || bytesRecv == WSAECONNRESET){ printf("Client: Connection Closed.\n"); break; }else printf("Client: recv() is OK.\n"); if (bytesRecv < 0) return 0; else printf("Client: Bytes received - %ld.\n", bytesRecv); } system("pause"); return 0; } server: int main(){ WORD wVersionRequested; WSADATA wsaData={0}; int wsaerr; // Using MAKEWORD macro, Winsock version request 2.2 wVersionRequested = MAKEWORD(2, 2); wsaerr = WSAStartup(wVersionRequested, &wsaData); if (wsaerr != 0){ /* Tell the user that we could not find a usable WinSock DLL.*/ printf("Server: The Winsock dll not found!\n"); return 0; }else{ printf("Server: The Winsock dll found!\n"); printf("Server: The status: %s.\n", wsaData.szSystemStatus); } /* Confirm that the WinSock DLL supports 2.2.*/ /* Note that if the DLL supports versions greater */ /* than 2.2 in addition to 2.2, it will still return */ /* 2.2 in wVersion since that is the version we */ /* requested. */ if (LOBYTE(wsaData.wVersion) != 2 || HIBYTE(wsaData.wVersion) != 2 ){ /* Tell the user that we could not find a usable WinSock DLL.*/ printf("Server: The dll do not support the Winsock version %u.%u!\n", LOBYTE(wsaData.wVersion), HIBYTE(wsaData.wVersion)); WSACleanup(); return 0; }else{ printf("Server: The dll supports the Winsock version %u.%u!\n", LOBYTE(wsaData.wVersion), HIBYTE(wsaData.wVersion)); printf("Server: The highest version this dll can support: %u.%u\n", LOBYTE(wsaData.wHighVersion), HIBYTE(wsaData.wHighVersion)); } //////////Create a socket//////////////////////// //Create a SOCKET object called m_socket. SOCKET m_socket; // Call the socket function and return its value to the m_socket variable. // For this application, use the Internet address family, streaming sockets, and the TCP/IP protocol. // using AF_INET family, TCP socket type and protocol of the AF_INET - IPv4 m_socket = socket(AF_INET, SOCK_STREAM, IPPROTO_TCP); // Check for errors to ensure that the socket is a valid socket. if (m_socket == INVALID_SOCKET){ printf("Server: Error at socket(): %ld\n", WSAGetLastError()); WSACleanup(); //return 0; }else{ printf("Server: socket() is OK!\n"); } ////////////////bind////////////////////////////// // Create a sockaddr_in object and set its values. sockaddr_in service; // AF_INET is the Internet address family. service.sin_family = AF_INET; // "127.0.0.1" is the local IP address to which the socket will be bound. service.sin_addr.s_addr = htons(INADDR_ANY);//inet_addr("127.0.0.1");//htons(INADDR_ANY); //inet_addr("192.168.1.2"); // 55555 is the port number to which the socket will be bound. // using the htons for big-endian service.sin_port = htons(12345); // Call the bind function, passing the created socket and the sockaddr_in structure as parameters. // Check for general errors. if (bind(m_socket, (SOCKADDR*)&service, sizeof(service)) == SOCKET_ERROR){ printf("Server: bind() failed: %ld.\n", WSAGetLastError()); closesocket(m_socket); //return 0; }else{ printf("Server: bind() is OK!\n"); } // Call the listen function, passing the created socket and the maximum number of allowed // connections to accept as parameters. Check for general errors. if (listen(m_socket, 1) == SOCKET_ERROR) printf("Server: listen(): Error listening on socket %ld.\n", WSAGetLastError()); else{ printf("Server: listen() is OK, I'm waiting for connections...\n"); } // Create a temporary SOCKET object called AcceptSocket for accepting connections. SOCKET AcceptSocket; // Create a continuous loop that checks for connections requests. If a connection // request occurs, call the accept function to handle the request. printf("Server: Waiting for a client to connect...\n"); printf("***Hint: Server is ready...run your client program...***\n"); // Do some verification... while (1){ AcceptSocket = SOCKET_ERROR; while (AcceptSocket == SOCKET_ERROR){ AcceptSocket = accept(m_socket, NULL, NULL); } // else, accept the connection... note: now it is wrong implementation !!!!!!!! !! !! (only 1 char) // When the client connection has been accepted, transfer control from the // temporary socket to the original socket and stop checking for new connections. printf("Server: Client Connected! Mammamija. \n"); m_socket = AcceptSocket; char recvBuf[200]=""; char * rc=recvBuf; int bytesRecv=recv(m_socket,recvBuf,64,0); if(bytesRecv==0 || bytesRecv==WSAECONNRESET){ cout<<"server: connection closed.\n"; }else{ cout<<"server: recv() is OK.\n"; if(bytesRecv<0){ return 0; }else{ printf("server: bytes received: %ld.\n",recvBuf); } }

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  • External USB attached drive works in Windows XP but not in Windows 7. How to fix?

    - by irrational John
    Earlier this week I purchased this "N52300 EZQuest Pro" external hard drive enclosure from here. I can connect the enclosure using USB 2.0 and access the files in both NTFS partitions on the MBR partitioned drive when I use either Windows XP (SP3) or Mac OS X 10.6. So it works as expected in XP & Snow Leopard. However, the enclosure does not work in Windows 7 (Home Premium) either 64-bit or 32-bit or in Ubuntu 10.04 (kernel 2.6.32-23-generic). I'm thinking this must be a Windows 7 driver problem because the enclosure works in XP & Snow Leopard. I do know that no special drivers are required to use this enclosure. It is supported using the USB mass storage drivers included with XP and OS X. It should also work fine using the mass storage support in Windows 7, no? FWIW, I have also tried using 32-bit Windows 7 on both my desktop, a Gigabyte GA-965P-DS3 with a Pentium Dual-Core E6500 @ 2.93GHz, and on my early 2008 MacBook. I see the same failure in both cases that I see with 64-bit Windows 7. So it doesn't appear to be specific to one hardware platform. I'm hoping someone out there can help me either get the enclosure to work in Windows 7 or convince me that the enclosure hardware is bad and should be RMAed. At the moment though an RMA seems pointless since this appears to be a (Windows 7) device driver problem. I have tried to track down any updates to the mass storage drivers included with Windows 7 but have so far come up empty. Heck, I can't even figure out how to place a bug report with Microsoft since apparently the grace period for Windows 7 email support is only a few months. I came across a link to some USB troubleshooting steps in another question. I haven't had a chance to look over the suggestions on that site or try them yet. Maybe tomorrow if I have time ... ;-) I'll finish up with some more details about the problem. When I connect the enclosure using USB to Windows 7 at first it appears everything worked. Windows detects the drive and installs a driver for it. Looking in Device Manager there is an entry under the Hard Drives section with the title, Hitachi HDT721010SLA360 USB Device. When you open Windows Disk Management the first time after the enclosure has been attached the drive appears as "Not initialize" and I'm prompted to initialize it. This is bogus. After all, the drive worked fine in XP so I know it has already been initialized, partitioned, and formatted. So of course I never try to initialize it "again". (It's a 1 GB drive and I don't want to lose the data on it). Except for this first time, the drive never shows up in Disk Management again unless I uninstall the Hitachi HDT721010SLA360 USB Device entry under Hard Drives, unplug, and then replug the enclosure. If I do that then the process in the previous paragraph repeats. In Ubuntu the enclosure never shows up at all at the file system level. Below are an excerpt from kern.log and an excerpt from the result of lsusb -v after attaching the enclosure. It appears that Ubuntu at first recongnizes the enclosure and is attempting to attach it, but encounters errors which prevent it from doing so. Unfortunately, I don't know whether any of this info is useful or not. excerpt from kern.log [ 2684.240015] usb 1-2: new high speed USB device using ehci_hcd and address 22 [ 2684.393618] usb 1-2: configuration #1 chosen from 1 choice [ 2684.395399] scsi17 : SCSI emulation for USB Mass Storage devices [ 2684.395570] usb-storage: device found at 22 [ 2684.395572] usb-storage: waiting for device to settle before scanning [ 2689.390412] usb-storage: device scan complete [ 2689.390894] scsi 17:0:0:0: Direct-Access Hitachi HDT721010SLA360 ST6O PQ: 0 ANSI: 4 [ 2689.392237] sd 17:0:0:0: Attached scsi generic sg7 type 0 [ 2689.395269] sd 17:0:0:0: [sde] 1953525168 512-byte logical blocks: (1.00 TB/931 GiB) [ 2689.395632] sd 17:0:0:0: [sde] Write Protect is off [ 2689.395636] sd 17:0:0:0: [sde] Mode Sense: 11 00 00 00 [ 2689.395639] sd 17:0:0:0: [sde] Assuming drive cache: write through [ 2689.412003] sd 17:0:0:0: [sde] Assuming drive cache: write through [ 2689.412009] sde: sde1 sde2 [ 2689.455759] sd 17:0:0:0: [sde] Assuming drive cache: write through [ 2689.455765] sd 17:0:0:0: [sde] Attached SCSI disk [ 2692.620017] usb 1-2: reset high speed USB device using ehci_hcd and address 22 [ 2707.740014] usb 1-2: device descriptor read/64, error -110 [ 2722.970103] usb 1-2: device descriptor read/64, error -110 [ 2723.200027] usb 1-2: reset high speed USB device using ehci_hcd and address 22 [ 2738.320019] usb 1-2: device descriptor read/64, error -110 [ 2753.550024] usb 1-2: device descriptor read/64, error -110 [ 2753.780020] usb 1-2: reset high speed USB device using ehci_hcd and address 22 [ 2758.810147] usb 1-2: device descriptor read/8, error -110 [ 2763.940142] usb 1-2: device descriptor read/8, error -110 [ 2764.170014] usb 1-2: reset high speed USB device using ehci_hcd and address 22 [ 2769.200141] usb 1-2: device descriptor read/8, error -110 [ 2774.330137] usb 1-2: device descriptor read/8, error -110 [ 2774.440069] usb 1-2: USB disconnect, address 22 [ 2774.440503] sd 17:0:0:0: Device offlined - not ready after error recovery [ 2774.590023] usb 1-2: new high speed USB device using ehci_hcd and address 23 [ 2789.710020] usb 1-2: device descriptor read/64, error -110 [ 2804.940020] usb 1-2: device descriptor read/64, error -110 [ 2805.170026] usb 1-2: new high speed USB device using ehci_hcd and address 24 [ 2820.290019] usb 1-2: device descriptor read/64, error -110 [ 2835.520027] usb 1-2: device descriptor read/64, error -110 [ 2835.750018] usb 1-2: new high speed USB device using ehci_hcd and address 25 [ 2840.780085] usb 1-2: device descriptor read/8, error -110 [ 2845.910079] usb 1-2: device descriptor read/8, error -110 [ 2846.140023] usb 1-2: new high speed USB device using ehci_hcd and address 26 [ 2851.170112] usb 1-2: device descriptor read/8, error -110 [ 2856.300077] usb 1-2: device descriptor read/8, error -110 [ 2856.410027] hub 1-0:1.0: unable to enumerate USB device on port 2 [ 2856.730033] usb 3-2: new full speed USB device using uhci_hcd and address 11 [ 2871.850017] usb 3-2: device descriptor read/64, error -110 [ 2887.080014] usb 3-2: device descriptor read/64, error -110 [ 2887.310011] usb 3-2: new full speed USB device using uhci_hcd and address 12 [ 2902.430021] usb 3-2: device descriptor read/64, error -110 [ 2917.660013] usb 3-2: device descriptor read/64, error -110 [ 2917.890016] usb 3-2: new full speed USB device using uhci_hcd and address 13 [ 2922.911623] usb 3-2: device descriptor read/8, error -110 [ 2928.051753] usb 3-2: device descriptor read/8, error -110 [ 2928.280013] usb 3-2: new full speed USB device using uhci_hcd and address 14 [ 2933.301876] usb 3-2: device descriptor read/8, error -110 [ 2938.431993] usb 3-2: device descriptor read/8, error -110 [ 2938.540073] hub 3-0:1.0: unable to enumerate USB device on port 2 excerpt from lsusb -v Bus 001 Device 017: ID 0dc4:0000 Macpower Peripherals, Ltd Device Descriptor: bLength 18 bDescriptorType 1 bcdUSB 2.00 bDeviceClass 0 (Defined at Interface level) bDeviceSubClass 0 bDeviceProtocol 0 bMaxPacketSize0 64 idVendor 0x0dc4 Macpower Peripherals, Ltd idProduct 0x0000 bcdDevice 0.01 iManufacturer 1 EZ QUEST iProduct 2 USB Mass Storage iSerial 3 220417 bNumConfigurations 1 Configuration Descriptor: bLength 9 bDescriptorType 2 wTotalLength 32 bNumInterfaces 1 bConfigurationValue 1 iConfiguration 5 Config0 bmAttributes 0xc0 Self Powered MaxPower 0mA Interface Descriptor: bLength 9 bDescriptorType 4 bInterfaceNumber 0 bAlternateSetting 0 bNumEndpoints 2 bInterfaceClass 8 Mass Storage bInterfaceSubClass 6 SCSI bInterfaceProtocol 80 Bulk (Zip) iInterface 4 Interface0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x01 EP 1 OUT bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0200 1x 512 bytes bInterval 0 Endpoint Descriptor: bLength 7 bDescriptorType 5 bEndpointAddress 0x81 EP 1 IN bmAttributes 2 Transfer Type Bulk Synch Type None Usage Type Data wMaxPacketSize 0x0200 1x 512 bytes bInterval 0 Device Qualifier (for other device speed): bLength 10 bDescriptorType 6 bcdUSB 2.00 bDeviceClass 0 (Defined at Interface level) bDeviceSubClass 0 bDeviceProtocol 0 bMaxPacketSize0 64 bNumConfigurations 1 Device Status: 0x0001 Self Powered Update: Results using Firewire to connect. Today I recieved a 1394b 9 pin to 1394a 6 pin cable which allowed me to connect the "EZQuest Pro" via Firewire. Everything works. When I use Firewire I can connect whether I'm using Windows 7 or Ubuntu 10.04. I even tried booting my Gigabyte desktop as an OS X 10.6.3 Hackintosh and it worked there as well. (Though if I recall correctly, it also worked when using USB 2.0 and booting OS X on the desktop. Certainly it works with USB 2.0 and my MacBook.) I believe the firmware on the device is at the latest level available, v1.07. I base this on the excerpt below from the OS X System Profiler which shows Firmware Revision: 0x107. Bottom line: It's nice that the enclosure is actually usable when I connect with Firewire. But I am still searching for an answer as to why it does not work correctly when using USB 2.0 in Windows 7 (and Ubuntu ... but really Windows 7 is my biggest concern). OXFORD IDE Device 1: Manufacturer: EZ QUEST Model: 0x0 GUID: 0x1D202E0220417 Maximum Speed: Up to 800 Mb/sec Connection Speed: Up to 400 Mb/sec Sub-units: OXFORD IDE Device 1 Unit: Unit Software Version: 0x10483 Unit Spec ID: 0x609E Firmware Revision: 0x107 Product Revision Level: ST6O Sub-units: OXFORD IDE Device 1 SBP-LUN: Capacity: 1 TB (1,000,204,886,016 bytes) Removable Media: Yes BSD Name: disk3 Partition Map Type: MBR (Master Boot Record) S.M.A.R.T. status: Not Supported

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  • Why did my flash drive become "read only" and (how) can I fix it?

    - by Bob
    I have a brand new flash drive (one week old) that has become marked as read only, by Windows, Kubuntu and a bootable partitioner. Why did this happen? Is it fixable? If it is, how can I fix this? The problem Firstly, this drive is new. It's certainly not been used enough to die from normal wear and tear, though I would not discount defective components. The drive itself has somehow become locked in a read only state. Windows' Disk management: Diskpart: Generic Flash Disk USB Device Disk ID: 33FA33FA Type : USB Status : Online Path : 0 Target : 0 LUN ID : 0 Location Path : UNAVAILABLE Current Read-only State : Yes Read-only : No Boot Disk : No Pagefile Disk : No Hibernation File Disk : No Crashdump Disk : No Clustered Disk : No What really confuses me is Current Read-only State : Yes and Read-only : No. Attempted solutions So far, I've tried: Formatting it in Windows (in Disk management, the format options are greyed out when right clicking). DiskPart Clean (CLEAN - Clear the configuration information, or all information, off the disk.): DISKPART> clean DiskPart has encountered an error: The media is write protected. See the System Event Log for more information. There was nothing in the event log. Windows command line format >format G: Insert new disk for drive G: and press ENTER when ready... The type of the file system is FAT32. Verifying 7740M Cannot format. This volume is write protected. Windows chkdsk: see below for details Kubuntu fsck (through VirtualBox USB passthrough): see below for details Acronis True Image to format, to convert to GPT, to destroy and rebuild MBR, basically anything: failed (could not write to MBR) Details (and a nice story) Background This was a brand new, generic, 8GB flash drive I wanted to create a multiboot flash drive with. It came formatted as FAT32, though oddly a little larger than most 8 GIGAbyte flash drives I've come across. Approximately 127MB was listed as "used" by Windows. I never discovered why. The end usable space was about what I normally expect from a 8GB drive (approx 7.4 GIBIbytes). I had thrown quite a few Linux distros on, along with a copy of Hiren's. They would all boot perfectly. They were put on with YUMI. When I tried to put the Knoppix DVD on, YUMI added an odd video option to its boot comman which caused Knoppix to boot with a black screen on X. ttys 1 through 6 still worked as text only interfaces. A few days later, I took some time to take that odd video option off, making the boot command match the one that comes with Knoppix. On the attempt to boot, Knoppix reported some form of LZMA corruption. Leading up to the current issue I was thinking the Knoppix files may have been corrupted somehow, so I tried reloading it. The drive was nearly full (45MB free), so I deleted a generic ISO that also was not booting. That went fine. I then went through YUMI to 'uninstall' Knoppix, i.e. delete files and remove from the menus. The files went first, then the menus were cleared successfully. However, the free space was stuck at about 700MB, same as it was before removing Knoppix. In the old Knoppix folder, there was a 0 byte file named KNOPPIX that could not be deleted. I tried reinserting the drive to delete this file - without safely removing, if that made a difference (hey, first time for everything). Running the standard Windows chkdsk scan without /r or /f reported errors found. Running with /r just got it stuck. I decided to give fsck a shot, so I loaded up my Kubuntu VM and attached the drive to it with VirtualBox's USB 2.0 passthrough. I umounted it (/dev/sda1) and ran a fsck. There are differences between boot sector and its backup. I chose No action. It told me FATs differ and asked me to select either the first or second FAT. Whichever I selected, I got a notice of Free cluster summary wrong. If I chose Correct, it gave a list of incorrect file names. To try to fix something, at least, I ran it with the -p option. Halfway through fixing the files, the VM froze - I ended its process about ten minutes later. Cause? My next attempt was to use YUMI, again, to rebuild the whole drive. I used YUMI's built in reformat (to FAT32) option and installed a Kubuntu ISO (700MB). The format was successful, however, the extract and copy of Kubuntu (which YUMI uses a 7zip binary for) froze at about 60% done. After waiting for about fifteen minutes (longer than the 3.5GB Knoppix ISO took last time), I pulled the drive out. The drive at this point was already formatted, SYSLINUX already installed, just waiting on the unpacking of an ISO and the modifying of the boot menus. Plugging it back in, it came up as normal - however, any write action would fail. Disk management reported it as read only. On reconnect, it would come up as normal but a write operation would cause it to go read only again. After a few attempts, it started coming up as read only on insertion. Attempts to fix This is when I ran through the attempts listed above, to try and reformat it in case of a faulty format. However the inability to do so even on a bootable disk indicated something more serious is wrong. chkdsk now reports nothing is wrong, and fsck still reports MBR inconsistencies, but now always chooses first FAT automatically after telling me FATs differ. It still does the same Free cluster summary wrong afterwards. I cannot run with -p anymore because it is now marked as read only. It also managed to corrupt my VM's disk somehow on the first attempt (yes, I'm sure I chose sda, which is mapped to a 7.4GB drive - I triple checked). Thank god for snapshots? I'm just about out of ideas. To my inexperienced mind it looks like something in the drive's firmware set it to read only "permanently" somehow - is there any way to reset this? I don't particularly care about keeping data, considering I've reformatted it twice. Also, fixes that keep me in Windows are better; it reduces the risk of me accidentally nuking my main hard drive. Update 1: I pulled apart the drive out of curiosity. As you can see, there are no obvious write protect switches. There is an IC on the other side, ALCOR branded labelled AU6989HL, if that matters. If there appears to be no way to fix this, I'll probably pull out the (glued down) card and put it in a card reader to check if it's the card or the controller that died. Update 2: I've pulled the card off, Windows detects the drive as a card reader now. The contacts on the card don't appear to be used, and there are several rows of holes on the card itself. Putting it into the card reader only detects about 30MB total, RAW. It's probably either the reader incorrectly reporting the card as faulty (as if a real SD card's write protect was switched on) or a bad contact somewhere. If nothing else, I have a spare 8GB Micro SD card now... as soon as I figure out how to format it as 8GB.

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  • ProFTPd server on Ubuntu getting access denied message when successfully authenticated?

    - by exxoid
    I have a Ubuntu box with a ProFTPD 1.3.4a Server, when I try to log in via my FTP Client I cannot do anything as it does not allow me to list directories; I have tried logging in as root and as a regular user and tried accessing different paths within the FTP Server. The error I get in my FTP Client is: Status: Retrieving directory listing... Command: CDUP Response: 250 CDUP command successful Command: PWD Response: 257 "/var" is the current directory Command: PASV Response: 227 Entering Passive Mode (172,16,4,22,237,205). Command: MLSD Response: 550 Access is denied. Error: Failed to retrieve directory listing Any idea? Here is the config of my proftpd: # # /etc/proftpd/proftpd.conf -- This is a basic ProFTPD configuration file. # To really apply changes, reload proftpd after modifications, if # it runs in daemon mode. It is not required in inetd/xinetd mode. # # Includes DSO modules Include /etc/proftpd/modules.conf # Set off to disable IPv6 support which is annoying on IPv4 only boxes. UseIPv6 off # If set on you can experience a longer connection delay in many cases. IdentLookups off ServerName "Drupal Intranet" ServerType standalone ServerIdent on "FTP Server ready" DeferWelcome on # Set the user and group that the server runs as User nobody Group nogroup MultilineRFC2228 on DefaultServer on ShowSymlinks on TimeoutNoTransfer 600 TimeoutStalled 600 TimeoutIdle 1200 DisplayLogin welcome.msg DisplayChdir .message true ListOptions "-l" DenyFilter \*.*/ # Use this to jail all users in their homes # DefaultRoot ~ # Users require a valid shell listed in /etc/shells to login. # Use this directive to release that constrain. # RequireValidShell off # Port 21 is the standard FTP port. Port 21 # In some cases you have to specify passive ports range to by-pass # firewall limitations. Ephemeral ports can be used for that, but # feel free to use a more narrow range. # PassivePorts 49152 65534 # If your host was NATted, this option is useful in order to # allow passive tranfers to work. You have to use your public # address and opening the passive ports used on your firewall as well. # MasqueradeAddress 1.2.3.4 # This is useful for masquerading address with dynamic IPs: # refresh any configured MasqueradeAddress directives every 8 hours <IfModule mod_dynmasq.c> # DynMasqRefresh 28800 </IfModule> # To prevent DoS attacks, set the maximum number of child processes # to 30. If you need to allow more than 30 concurrent connections # at once, simply increase this value. Note that this ONLY works # in standalone mode, in inetd mode you should use an inetd server # that allows you to limit maximum number of processes per service # (such as xinetd) MaxInstances 30 # Set the user and group that the server normally runs at. # Umask 022 is a good standard umask to prevent new files and dirs # (second parm) from being group and world writable. Umask 022 022 # Normally, we want files to be overwriteable. AllowOverwrite on # Uncomment this if you are using NIS or LDAP via NSS to retrieve passwords: # PersistentPasswd off # This is required to use both PAM-based authentication and local passwords AuthPAMConfig proftpd AuthOrder mod_auth_pam.c* mod_auth_unix.c # Be warned: use of this directive impacts CPU average load! # Uncomment this if you like to see progress and transfer rate with ftpwho # in downloads. That is not needed for uploads rates. # UseSendFile off TransferLog /var/log/proftpd/xferlog SystemLog /var/log/proftpd/proftpd.log # Logging onto /var/log/lastlog is enabled but set to off by default #UseLastlog on # In order to keep log file dates consistent after chroot, use timezone info # from /etc/localtime. If this is not set, and proftpd is configured to # chroot (e.g. DefaultRoot or <Anonymous>), it will use the non-daylight # savings timezone regardless of whether DST is in effect. #SetEnv TZ :/etc/localtime <IfModule mod_quotatab.c> QuotaEngine off </IfModule> <IfModule mod_ratio.c> Ratios off </IfModule> # Delay engine reduces impact of the so-called Timing Attack described in # http://www.securityfocus.com/bid/11430/discuss # It is on by default. <IfModule mod_delay.c> DelayEngine on </IfModule> <IfModule mod_ctrls.c> ControlsEngine off ControlsMaxClients 2 ControlsLog /var/log/proftpd/controls.log ControlsInterval 5 ControlsSocket /var/run/proftpd/proftpd.sock </IfModule> <IfModule mod_ctrls_admin.c> AdminControlsEngine off </IfModule> # # Alternative authentication frameworks # #Include /etc/proftpd/ldap.conf #Include /etc/proftpd/sql.conf # # This is used for FTPS connections # #Include /etc/proftpd/tls.conf # # Useful to keep VirtualHost/VirtualRoot directives separated # #Include /etc/proftpd/virtuals.con # A basic anonymous configuration, no upload directories. # <Anonymous ~ftp> # User ftp # Group nogroup # # We want clients to be able to login with "anonymous" as well as "ftp" # UserAlias anonymous ftp # # Cosmetic changes, all files belongs to ftp user # DirFakeUser on ftp # DirFakeGroup on ftp # # RequireValidShell off # # # Limit the maximum number of anonymous logins # MaxClients 10 # # # We want 'welcome.msg' displayed at login, and '.message' displayed # # in each newly chdired directory. # DisplayLogin welcome.msg # DisplayChdir .message # # # Limit WRITE everywhere in the anonymous chroot # <Directory *> # <Limit WRITE> # DenyAll # </Limit> # </Directory> # # # Uncomment this if you're brave. # # <Directory incoming> # # # Umask 022 is a good standard umask to prevent new files and dirs # # # (second parm) from being group and world writable. # # Umask 022 022 # # <Limit READ WRITE> # # DenyAll # # </Limit> # # <Limit STOR> # # AllowAll # # </Limit> # # </Directory> # # </Anonymous> # Include other custom configuration files Include /etc/proftpd/conf.d/ UseReverseDNS off <Global> RootLogin on UseFtpUsers on ServerIdent on DefaultChdir /var/www DeleteAbortedStores on LoginPasswordPrompt on AccessGrantMsg "You have been authenticated successfully." </Global> Any idea what could be wrong? Thanks for your help!

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  • Integrating HTML into Silverlight Applications

    - by dwahlin
    Looking for a way to display HTML content within a Silverlight application? If you haven’t tried doing that before it can be challenging at first until you know a few tricks of the trade.  Being able to display HTML is especially handy when you’re required to display RSS feeds (with embedded HTML), SQL Server Reporting Services reports, PDF files (not actually HTML – but the techniques discussed will work), or other HTML content.  In this post I'll discuss three options for displaying HTML content in Silverlight applications and describe how my company is using these techniques in client applications. Displaying HTML Overlays If you need to display HTML over a Silverlight application (such as an RSS feed containing HTML data in it) you’ll need to set the Silverlight control’s windowless parameter to true. This can be done using the object tag as shown next: <object data="data:application/x-silverlight-2," type="application/x-silverlight-2" width="100%" height="100%"> <param name="source" value="ClientBin/HTMLAndSilverlight.xap"/> <param name="onError" value="onSilverlightError" /> <param name="background" value="white" /> <param name="minRuntimeVersion" value="4.0.50401.0" /> <param name="autoUpgrade" value="true" /> <param name="windowless" value="true" /> <a href="http://go.microsoft.com/fwlink/?LinkID=149156&v=4.0.50401.0" style="text-decoration:none"> <img src="http://go.microsoft.com/fwlink/?LinkId=161376" alt="Get Microsoft Silverlight" style="border-style:none"/> </a> </object> By setting the control to “windowless” you can overlay HTML objects by using absolute positioning and other CSS techniques. Keep in mind that on Windows machines the windowless setting can result in a performance hit when complex animations or HD video are running since the plug-in content is displayed directly by the browser window. It goes without saying that you should only set windowless to true when you really need the functionality it offers. For example, if I want to display my blog’s RSS content on top of a Silverlight application I could set windowless to true and create a user control that grabbed the content and output it using a DataList control: <style type="text/css"> a {text-decoration:none;font-weight:bold;font-size:14pt;} </style> <div style="margin-top:10px; margin-left:10px;margin-right:5px;"> <asp:DataList ID="RSSDataList" runat="server" DataSourceID="RSSDataSource"> <ItemTemplate> <a href='<%# XPath("link") %>'><%# XPath("title") %></a> <br /> <%# XPath("description") %> <br /> </ItemTemplate> </asp:DataList> <asp:XmlDataSource ID="RSSDataSource" DataFile="http://weblogs.asp.net/dwahlin/rss.aspx" XPath="rss/channel/item" CacheDuration="60" runat="server" /> </div> The user control can then be placed in the page hosting the Silverlight control as shown below. This example adds a Close button, additional content to display in the overlay window and the HTML generated from the user control. <div id="RSSDiv"> <div style="background-color:#484848;border:1px solid black;height:35px;width:100%;"> <img alt="Close Button" align="right" src="Images/Close.png" onclick="HideOverlay();" style="cursor:pointer;" /> </div> <div style="overflow:auto;width:800px;height:565px;"> <div style="float:left;width:100px;height:103px;margin-left:10px;margin-top:5px;"> <img src="http://weblogs.asp.net/blogs/dwahlin/dan2008.jpg" style="border:1px solid Gray" /> </div> <div style="float:left;width:300px;height:103px;margin-top:5px;"> <a href="http://weblogs.asp.net/dwahlin" style="margin-left:10px;font-size:20pt;">Dan Wahlin's Blog</a> </div> <br /><br /><br /> <div style="clear:both;margin-top:20px;"> <uc:BlogRoller ID="BlogRoller" runat="server" /> </div> </div> </div> Of course, we wouldn’t want the RSS HTML content to be shown until requested. Once it’s requested the absolute position of where it should show above the Silverlight control can be set using standard CSS styles. The following ID selector named #RSSDiv handles hiding the overlay div shown above and determines where it will be display on the screen. #RSSDiv { background-color:White; position:absolute; top:100px; left:300px; width:800px; height:600px; border:1px solid black; display:none; } Now that the HTML content to display above the Silverlight control is set, how can we show it as a user clicks a HyperlinkButton or other control in the application? Fortunately, Silverlight provides an excellent HTML bridge that allows direct access to content hosted within a page. The following code shows two JavaScript functions that can be called from Siverlight to handle showing or hiding HTML overlay content. The two functions rely on jQuery (http://www.jQuery.com) to make it easy to select HTML objects and manipulate their properties: function ShowOverlay() { rssDiv.css('display', 'block'); } function HideOverlay() { rssDiv.css('display', 'none'); } Calling the ShowOverlay function is as simple as adding the following code into the Silverlight application within a button’s Click event handler: private void OverlayHyperlinkButton_Click(object sender, RoutedEventArgs e) { HtmlPage.Window.Invoke("ShowOverlay"); } The result of setting the Silverlight control’s windowless parameter to true and showing the HTML overlay content is shown in the following screenshot:   Thinking Outside the Box to Show HTML Content Setting the windowless parameter to true may not be a viable option for some Silverlight applications or you may simply want to go about showing HTML content a different way. The next technique I’ll show takes advantage of simple HTML, CSS and JavaScript code to handle showing HTML content while a Silverlight application is running in the browser. Keep in mind that with Silverlight’s HTML bridge feature you can always pop-up HTML content in a new browser window using code similar to the following: System.Windows.Browser.HtmlPage.Window.Navigate( new Uri("http://silverlight.net"), "_blank"); For this example I’ll demonstrate how to hide the Silverlight application while maximizing a container div containing the HTML content to show. This allows HTML content to take up the full screen area of the browser without having to set windowless to true and when done right can make the user feel like they never left the Silverlight application. The following HTML shows several div elements that are used to display HTML within the same browser window as the Silverlight application: <div id="JobPlanDiv"> <div style="vertical-align:middle"> <img alt="Close Button" align="right" src="Images/Close.png" onclick="HideJobPlanIFrame();" style="cursor:pointer;" /> </div> <div id="JobPlan_IFrame_Container" style="height:95%;width:100%;margin-top:37px;"></div> </div> The JobPlanDiv element acts as a container for two other divs that handle showing a close button and hosting an iframe that will be added dynamically at runtime. JobPlanDiv isn’t visible when the Silverlight application loads due to the following ID selector added into the page: #JobPlanDiv { position:absolute; background-color:#484848; overflow:hidden; left:0; top:0; height:100%; width:100%; display:none; } When the HTML content needs to be shown or hidden the JavaScript functions shown next can be used: var jobPlanIFrameID = 'JobPlan_IFrame'; var slHost = null; var jobPlanContainer = null; var jobPlanIFrameContainer = null; var rssDiv = null; $(document).ready(function () { slHost = $('#silverlightControlHost'); jobPlanContainer = $('#JobPlanDiv'); jobPlanIFrameContainer = $('#JobPlan_IFrame_Container'); rssDiv = $('#RSSDiv'); }); function ShowJobPlanIFrame(url) { jobPlanContainer.css('display', 'block'); $('<iframe id="' + jobPlanIFrameID + '" src="' + url + '" style="height:100%;width:100%;" />') .appendTo(jobPlanIFrameContainer); slHost.css('width', '0%'); } function HideJobPlanIFrame() { jobPlanContainer.css('display', 'none'); $('#' + jobPlanIFrameID).remove(); slHost.css('width', '100%'); } ShowJobPlanIFrame() handles showing the JobPlanDiv div and adding an iframe into it dynamically. Once JobPlanDiv is shown, the Silverlight control host has its width set to a value of 0% to allow the control to stay alive while making it invisible to the user. I found that this technique works better across multiple browsers as opposed to manipulating the Silverlight control host div’s display or visibility properties. Now that you’ve seen the code to handle showing and hiding the HTML content area, let’s switch focus to the Silverlight application. As a user clicks on a link such as “View Report” the ShowJobPlanIFrame() JavaScript function needs to be called. The following code handles that task: private void ReportHyperlinkButton_Click(object sender, RoutedEventArgs e) { ShowBrowser(_BaseUrl + "/Report.aspx"); } public void ShowBrowser(string url) { HtmlPage.Window.Invoke("ShowJobPlanIFrame", url); } Any URL can be passed into the ShowBrowser() method which handles invoking the JavaScript function. This includes standard web pages or even PDF files. We’ve used this technique frequently with our SmartPrint control (http://www.smartwebcontrols.com) which converts Silverlight screens into PDF documents and displays them. Here’s an example of the content generated:   Silverlight 4’s WebBrowser Control Both techniques shown to this point work well when Silverlight is running in-browser but not so well when it’s running out-of-browser since there’s no host page that you can access using the HTML bridge. Fortunately, Silverlight 4 provides a WebBrowser control that can be used to perform the same functionality quite easily. We’re currently using it in client applications to display PDF documents, SSRS reports and standard HTML content. Using the WebBrowser control simplifies the application quite a bit since no JavaScript is required if the application only runs out-of-browser. Here’s a simple example of defining the WebBrowser control in XAML. I typically define it in MainPage.xaml when a Silverlight Navigation template is used to create the project so that I can re-use the functionality across multiple screens. <Grid x:Name="WebBrowserGrid" HorizontalAlignment="Stretch" VerticalAlignment="Stretch" Visibility="Collapsed"> <StackPanel HorizontalAlignment="Stretch" VerticalAlignment="Stretch"> <Border Background="#484848" HorizontalAlignment="Stretch" Height="40"> <Image x:Name="WebBrowserImage" Width="100" Height="33" Cursor="Hand" HorizontalAlignment="Right" Source="/HTMLAndSilverlight;component/Assets/Images/Close.png" MouseLeftButtonDown="WebBrowserImage_MouseLeftButtonDown" /> </Border> <WebBrowser x:Name="JobPlanReportWebBrowser" HorizontalAlignment="Stretch" VerticalAlignment="Stretch" /> </StackPanel> </Grid> Looking through the XAML you can see that a close image is defined along with the WebBrowser control. Because the URL that the WebBrowser should navigate to isn’t known at design time no value is assigned to the control’s Source property. If the XAML shown above is left “as is” you’ll find that any HTML content assigned to the WebBrowser doesn’t display properly. This is due to no height or width being set on the control. To handle this issue the following code is added into the XAML’s code-behind file to dynamically determine the height and width of the page and assign it to the WebBrowser. This is done by handling the SizeChanged event. void MainPage_SizeChanged(object sender, SizeChangedEventArgs e) { WebBrowserGrid.Height = JobPlanReportWebBrowser.Height = ActualHeight; WebBrowserGrid.Width = JobPlanReportWebBrowser.Width = ActualWidth; } When the user wants to view HTML content they click a button which executes the code shown in next: public void ShowBrowser(string url) { if (Application.Current.IsRunningOutOfBrowser) { JobPlanReportWebBrowser.NavigateToString("<html><body><iframe src='" + url + "' style='width:100%;height:97%;' /></body></html>"); WebBrowserGrid.Visibility = Visibility.Visible; } else { HtmlPage.Window.Invoke("ShowJobPlanIFrame", url); } } private void WebBrowserImage_MouseLeftButtonDown(object sender, MouseButtonEventArgs e) { WebBrowserGrid.Visibility = Visibility.Collapsed; }   Looking through the code you’ll see that it checks to see if the Silverlight application is running out-of-browser and then either displays the WebBrowser control or runs the JavaScript function discussed earlier. Although the WebBrowser control’s Source property could be assigned the URI of the page to navigate to, by assigning HTML content using the NavigateToString() method and adding an iframe, content can be shown from any site including cross-domain sites. This is especially handy when you need to grab a page from a reporting site that’s in a different domain than the Silverlight application. Here’s an example of viewing  PDF file inside of an out-of-browser application. The first image shows the application running out-of-browser before the user clicks a PDF HyperlinkButton.  The second image shows the PDF being displayed.   While there are certainly other techniques that can be used, the ones shown here have worked well for us in different applications and provide the ability to display HTML content in-browser or out-of-browser. Feel free to add a comment if you have another tip or trick you like to use when working with HTML content in Silverlight applications.   Download Code Sample   For more information about onsite, online and video training, mentoring and consulting solutions for .NET, SharePoint or Silverlight please visit http://www.thewahlingroup.com.

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  • WCF WS-Security and WSE Nonce Authentication

    - by Rick Strahl
    WCF makes it fairly easy to access WS-* Web Services, except when you run into a service format that it doesn't support. Even then WCF provides a huge amount of flexibility to make the service clients work, however finding the proper interfaces to make that happen is not easy to discover and for the most part undocumented unless you're lucky enough to run into a blog, forum or StackOverflow post on the matter. This is definitely true for the Password Nonce as part of the WS-Security/WSE protocol, which is not natively supported in WCF. Specifically I had a need to create a WCF message on the client that includes a WS-Security header that looks like this from their spec document:<soapenv:Header> <wsse:Security soapenv:mustUnderstand="1" xmlns:wsse="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"> <wsse:UsernameToken wsu:Id="UsernameToken-8" xmlns:wsu="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd"> <wsse:Username>TeStUsErNaMe1</wsse:Username> <wsse:Password Type="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#PasswordText" >TeStPaSsWoRd1</wsse:Password> <wsse:Nonce EncodingType="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-soap-message-security-1.0#Base64Binary" >f8nUe3YupTU5ISdCy3X9Gg==</wsse:Nonce> <wsu:Created>2011-05-04T19:01:40.981Z</wsu:Created> </wsse:UsernameToken> </wsse:Security> </soapenv:Header> Specifically, the Nonce and Created keys are what WCF doesn't create or have a built in formatting for. Why is there a nonce? My first thought here was WTF? The username and password are there in clear text, what does the Nonce accomplish? The Nonce and created keys are are part of WSE Security specification and are meant to allow the server to detect and prevent replay attacks. The hashed nonce should be unique per request which the server can store and check for before running another request thus ensuring that a request is not replayed with exactly the same values. Basic ServiceUtl Import - not much Luck The first thing I did when I imported this service with a service reference was to simply import it as a Service Reference. The Add Service Reference import automatically detects that WS-Security is required and appropariately adds the WS-Security to the basicHttpBinding in the config file:<?xml version="1.0" encoding="utf-8" ?> <configuration> <system.serviceModel> <bindings> <basicHttpBinding> <binding name="RealTimeOnlineSoapBinding"> <security mode="Transport" /> </binding> <binding name="RealTimeOnlineSoapBinding1" /> </basicHttpBinding> </bindings> <client> <endpoint address="https://notarealurl.com:443/services/RealTimeOnline" binding="basicHttpBinding" bindingConfiguration="RealTimeOnlineSoapBinding" contract="RealTimeOnline.RealTimeOnline" name="RealTimeOnline" /> </client> </system.serviceModel> </configuration> If if I run this as is using code like this:var client = new RealTimeOnlineClient(); client.ClientCredentials.UserName.UserName = "TheUsername"; client.ClientCredentials.UserName.Password = "ThePassword"; … I get nothing in terms of WS-Security headers. The request is sent, but the the binding expects transport level security to be applied, rather than message level security. To fix this so that a WS-Security message header is sent the security mode can be changed to: <security mode="TransportWithMessageCredential" /> Now if I re-run I at least get a WS-Security header which looks like this:<s:Envelope xmlns:s="http://schemas.xmlsoap.org/soap/envelope/" xmlns:u="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd"> <s:Header> <o:Security s:mustUnderstand="1" xmlns:o="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"> <u:Timestamp u:Id="_0"> <u:Created>2012-11-24T02:55:18.011Z</u:Created> <u:Expires>2012-11-24T03:00:18.011Z</u:Expires> </u:Timestamp> <o:UsernameToken u:Id="uuid-18c215d4-1106-40a5-8dd1-c81fdddf19d3-1"> <o:Username>TheUserName</o:Username> <o:Password Type="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#PasswordText" >ThePassword</o:Password> </o:UsernameToken> </o:Security> </s:Header> Closer! Now the WS-Security header is there along with a timestamp field (which might not be accepted by some WS-Security expecting services), but there's no Nonce or created timestamp as required by my original service. Using a CustomBinding instead My next try was to go with a CustomBinding instead of basicHttpBinding as it allows a bit more control over the protocol and transport configurations for the binding. Specifically I can explicitly specify the message protocol(s) used. Using configuration file settings here's what the config file looks like:<?xml version="1.0"?> <configuration> <system.serviceModel> <bindings> <customBinding> <binding name="CustomSoapBinding"> <security includeTimestamp="false" authenticationMode="UserNameOverTransport" defaultAlgorithmSuite="Basic256" requireDerivedKeys="false" messageSecurityVersion="WSSecurity10WSTrustFebruary2005WSSecureConversationFebruary2005WSSecurityPolicy11BasicSecurityProfile10"> </security> <textMessageEncoding messageVersion="Soap11"></textMessageEncoding> <httpsTransport maxReceivedMessageSize="2000000000"/> </binding> </customBinding> </bindings> <client> <endpoint address="https://notrealurl.com:443/services/RealTimeOnline" binding="customBinding" bindingConfiguration="CustomSoapBinding" contract="RealTimeOnline.RealTimeOnline" name="RealTimeOnline" /> </client> </system.serviceModel> <startup> <supportedRuntime version="v4.0" sku=".NETFramework,Version=v4.0"/> </startup> </configuration> This ends up creating a cleaner header that's missing the timestamp field which can cause some services problems. The WS-Security header output generated with the above looks like this:<s:Header> <o:Security s:mustUnderstand="1" xmlns:o="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"> <o:UsernameToken u:Id="uuid-291622ca-4c11-460f-9886-ac1c78813b24-1"> <o:Username>TheUsername</o:Username> <o:Password Type="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#PasswordText" >ThePassword</o:Password> </o:UsernameToken> </o:Security> </s:Header> This is closer as it includes only the username and password. The key here is the protocol for WS-Security:messageSecurityVersion="WSSecurity10WSTrustFebruary2005WSSecureConversationFebruary2005WSSecurityPolicy11BasicSecurityProfile10" which explicitly specifies the protocol version. There are several variants of this specification but none of them seem to support the nonce unfortunately. This protocol does allow for optional omission of the Nonce and created timestamp provided (which effectively makes those keys optional). With some services I tried that requested a Nonce just using this protocol actually worked where the default basicHttpBinding failed to connect, so this is a possible solution for access to some services. Unfortunately for my target service that was not an option. The nonce has to be there. Creating Custom ClientCredentials As it turns out WCF doesn't have support for the Digest Nonce as part of WS-Security, and so as far as I can tell there's no way to do it just with configuration settings. I did a bunch of research on this trying to find workarounds for this, and I did find a couple of entries on StackOverflow as well as on the MSDN forums. However, none of these are particularily clear and I ended up using bits and pieces of several of them to arrive at a working solution in the end. http://stackoverflow.com/questions/896901/wcf-adding-nonce-to-usernametoken http://social.msdn.microsoft.com/Forums/en-US/wcf/thread/4df3354f-0627-42d9-b5fb-6e880b60f8ee The latter forum message is the more useful of the two (the last message on the thread in particular) and it has most of the information required to make this work. But it took some experimentation for me to get this right so I'll recount the process here maybe a bit more comprehensively. In order for this to work a number of classes have to be overridden: ClientCredentials ClientCredentialsSecurityTokenManager WSSecurityTokenizer The idea is that we need to create a custom ClientCredential class to hold the custom properties so they can be set from the UI or via configuration settings. The TokenManager and Tokenizer are mainly required to allow the custom credentials class to flow through the WCF pipeline and eventually provide custom serialization. Here are the three classes required and their full implementations:public class CustomCredentials : ClientCredentials { public CustomCredentials() { } protected CustomCredentials(CustomCredentials cc) : base(cc) { } public override System.IdentityModel.Selectors.SecurityTokenManager CreateSecurityTokenManager() { return new CustomSecurityTokenManager(this); } protected override ClientCredentials CloneCore() { return new CustomCredentials(this); } } public class CustomSecurityTokenManager : ClientCredentialsSecurityTokenManager { public CustomSecurityTokenManager(CustomCredentials cred) : base(cred) { } public override System.IdentityModel.Selectors.SecurityTokenSerializer CreateSecurityTokenSerializer(System.IdentityModel.Selectors.SecurityTokenVersion version) { return new CustomTokenSerializer(System.ServiceModel.Security.SecurityVersion.WSSecurity11); } } public class CustomTokenSerializer : WSSecurityTokenSerializer { public CustomTokenSerializer(SecurityVersion sv) : base(sv) { } protected override void WriteTokenCore(System.Xml.XmlWriter writer, System.IdentityModel.Tokens.SecurityToken token) { UserNameSecurityToken userToken = token as UserNameSecurityToken; string tokennamespace = "o"; DateTime created = DateTime.Now; string createdStr = created.ToString("yyyy-MM-ddThh:mm:ss.fffZ"); // unique Nonce value - encode with SHA-1 for 'randomness' // in theory the nonce could just be the GUID by itself string phrase = Guid.NewGuid().ToString(); var nonce = GetSHA1String(phrase); // in this case password is plain text // for digest mode password needs to be encoded as: // PasswordAsDigest = Base64(SHA-1(Nonce + Created + Password)) // and profile needs to change to //string password = GetSHA1String(nonce + createdStr + userToken.Password); string password = userToken.Password; writer.WriteRaw(string.Format( "<{0}:UsernameToken u:Id=\"" + token.Id + "\" xmlns:u=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd\">" + "<{0}:Username>" + userToken.UserName + "</{0}:Username>" + "<{0}:Password Type=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#PasswordText\">" + password + "</{0}:Password>" + "<{0}:Nonce EncodingType=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-soap-message-security-1.0#Base64Binary\">" + nonce + "</{0}:Nonce>" + "<u:Created>" + createdStr + "</u:Created></{0}:UsernameToken>", tokennamespace)); } protected string GetSHA1String(string phrase) { SHA1CryptoServiceProvider sha1Hasher = new SHA1CryptoServiceProvider(); byte[] hashedDataBytes = sha1Hasher.ComputeHash(Encoding.UTF8.GetBytes(phrase)); return Convert.ToBase64String(hashedDataBytes); } } Realistically only the CustomTokenSerializer has any significant code in. The code there deals with actually serializing the custom credentials using low level XML semantics by writing output into an XML writer. I can't take credit for this code - most of the code comes from the MSDN forum post mentioned earlier - I made a few adjustments to simplify the nonce generation and also added some notes to allow for PasswordDigest generation. Per spec the nonce is nothing more than a unique value that's supposed to be 'random'. I'm thinking that this value can be any string that's unique and a GUID on its own probably would have sufficed. Comments on other posts that GUIDs can be potentially guessed are highly exaggerated to say the least IMHO. To satisfy even that aspect though I added the SHA1 encryption and binary decoding to give a more random value that would be impossible to 'guess'. The original example from the forum post used another level of encoding and decoding to string in between - but that really didn't accomplish anything but extra overhead. The header output generated from this looks like this:<s:Header> <o:Security s:mustUnderstand="1" xmlns:o="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"> <o:UsernameToken u:Id="uuid-f43d8b0d-0ebb-482e-998d-f544401a3c91-1" xmlns:u="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd"> <o:Username>TheUsername</o:Username> <o:Password Type="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#PasswordText">ThePassword</o:Password> <o:Nonce EncodingType="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-soap-message-security-1.0#Base64Binary" >PjVE24TC6HtdAnsf3U9c5WMsECY=</o:Nonce> <u:Created>2012-11-23T07:10:04.670Z</u:Created> </o:UsernameToken> </o:Security> </s:Header> which is exactly as it should be. Password Digest? In my case the password is passed in plain text over an SSL connection, so there's no digest required so I was done with the code above. Since I don't have a service handy that requires a password digest,  I had no way of testing the code for the digest implementation, but here is how this is likely to work. If you need to pass a digest encoded password things are a little bit trickier. The password type namespace needs to change to: http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#Digest and then the password value needs to be encoded. The format for password digest encoding is this: Base64(SHA-1(Nonce + Created + Password)) and it can be handled in the code above with this code (that's commented in the snippet above): string password = GetSHA1String(nonce + createdStr + userToken.Password); The entire WriteTokenCore method for digest code looks like this:protected override void WriteTokenCore(System.Xml.XmlWriter writer, System.IdentityModel.Tokens.SecurityToken token) { UserNameSecurityToken userToken = token as UserNameSecurityToken; string tokennamespace = "o"; DateTime created = DateTime.Now; string createdStr = created.ToString("yyyy-MM-ddThh:mm:ss.fffZ"); // unique Nonce value - encode with SHA-1 for 'randomness' // in theory the nonce could just be the GUID by itself string phrase = Guid.NewGuid().ToString(); var nonce = GetSHA1String(phrase); string password = GetSHA1String(nonce + createdStr + userToken.Password); writer.WriteRaw(string.Format( "<{0}:UsernameToken u:Id=\"" + token.Id + "\" xmlns:u=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd\">" + "<{0}:Username>" + userToken.UserName + "</{0}:Username>" + "<{0}:Password Type=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-username-token-profile-1.0#Digest\">" + password + "</{0}:Password>" + "<{0}:Nonce EncodingType=\"http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-soap-message-security-1.0#Base64Binary\">" + nonce + "</{0}:Nonce>" + "<u:Created>" + createdStr + "</u:Created></{0}:UsernameToken>", tokennamespace)); } I had no service to connect to to try out Digest auth - if you end up needing it and get it to work please drop a comment… How to use the custom Credentials The easiest way to use the custom credentials is to create the client in code. Here's a factory method I use to create an instance of my service client:  public static RealTimeOnlineClient CreateRealTimeOnlineProxy(string url, string username, string password) { if (string.IsNullOrEmpty(url)) url = "https://notrealurl.com:443/cows/services/RealTimeOnline"; CustomBinding binding = new CustomBinding(); var security = TransportSecurityBindingElement.CreateUserNameOverTransportBindingElement(); security.IncludeTimestamp = false; security.DefaultAlgorithmSuite = SecurityAlgorithmSuite.Basic256; security.MessageSecurityVersion = MessageSecurityVersion.WSSecurity10WSTrustFebruary2005WSSecureConversationFebruary2005WSSecurityPolicy11BasicSecurityProfile10; var encoding = new TextMessageEncodingBindingElement(); encoding.MessageVersion = MessageVersion.Soap11; var transport = new HttpsTransportBindingElement(); transport.MaxReceivedMessageSize = 20000000; // 20 megs binding.Elements.Add(security); binding.Elements.Add(encoding); binding.Elements.Add(transport); RealTimeOnlineClient client = new RealTimeOnlineClient(binding, new EndpointAddress(url)); // to use full client credential with Nonce uncomment this code: // it looks like this might not be required - the service seems to work without it client.ChannelFactory.Endpoint.Behaviors.Remove<System.ServiceModel.Description.ClientCredentials>(); client.ChannelFactory.Endpoint.Behaviors.Add(new CustomCredentials()); client.ClientCredentials.UserName.UserName = username; client.ClientCredentials.UserName.Password = password; return client; } This returns a service client that's ready to call other service methods. The key item in this code is the ChannelFactory endpoint behavior modification that that first removes the original ClientCredentials and then adds the new one. The ClientCredentials property on the client is read only and this is the way it has to be added.   Summary It's a bummer that WCF doesn't suport WSE Security authentication with nonce values out of the box. From reading the comments in posts/articles while I was trying to find a solution, I found that this feature was omitted by design as this protocol is considered unsecure. While I agree that plain text passwords are rarely a good idea even if they go over secured SSL connection as WSE Security does, there are unfortunately quite a few services (mosly Java services I suspect) that use this protocol. I've run into this twice now and trying to find a solution online I can see that this is not an isolated problem - many others seem to have struggled with this. It seems there are about a dozen questions about this on StackOverflow all with varying incomplete answers. Hopefully this post provides a little more coherent content in one place. Again I marvel at WCF and its breadth of support for protocol features it has in a single tool. And even when it can't handle something there are ways to get it working via extensibility. But at the same time I marvel at how freaking difficult it is to arrive at these solutions. I mean there's no way I could have ever figured this out on my own. It takes somebody working on the WCF team or at least being very, very intricately involved in the innards of WCF to figure out the interconnection of the various objects to do this from scratch. Luckily this is an older problem that has been discussed extensively online and I was able to cobble together a solution from the online content. I'm glad it worked out that way, but it feels dirty and incomplete in that there's a whole learning path that was omitted to get here… Man am I glad I'm not dealing with SOAP services much anymore. REST service security - even when using some sort of federation is a piece of cake by comparison :-) I'm sure once standards bodies gets involved we'll be right back in security standard hell…© Rick Strahl, West Wind Technologies, 2005-2012Posted in WCF  Web Services   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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  • IIS SSL Certificate Renewal Pain

    - by Rick Strahl
    I’m in the middle of my annual certificate renewal for the West Wind site and I can honestly say that I hate IIS’s certificate system.  When it works it’s fine, but when it doesn’t man can it be a pain. Because I deal with public certificates on my site merely once a year, and you have to perform the certificate dance just the right way, I seem to run into some sort of trouble every year, thinking that Microsoft surely must have addressed the issues I ran into previously – HA! Not so. Don’t ever use the Renew Certificate Feature in IIS! The first rule that I should have never forgotten is that certificate renewals in IIS (7 is what I’m using but I think it’s no different in 7.5 and 8), simply don’t work if you’re submitting to get a public certificate from a certificate authority. I use DNSimple for my DNS domain management and SSL certificates because they provide ridiculously easy domain management and good prices for SSL certs – especially wildcard certificates, which is what I use on west-wind.com. Certificates in IIS can be found pegged to the machine root. If you go into the IIS Manager, go to the machine root the tree and then click on certificates and you then get various certificate options: Both of these options create a new Certificate request (CSR), which is just a text file. But if you’re silly enough like me to click on the Renew button on your old certificate, you’ll find that you end up generating a very long Certificate Request that looks nothing like the original certificate request and the format that’s used for this is not accepted by most certificate authorities. While I’m not sure exactly what the problem is, it simply looks like IIS is respecting none of your original certificate bit size choices and is generating a huge certificate request that is 3 times the size of a ‘normal’ certificate request. The end result is (and I’ve done this at least twice now) is that the certificate processor is likely to fail processing those renewals. Always create a new Certificate While it’s a little more work and you have to remember how to fill out the certificate request properly, this is the safe way to make sure your certificate generates properly. First comes the Distinguished Name Properties dialog: Ah yes you have to love the nomenclature of this stuff. Distinguished name, Common name – WTF is a common name? It doesn’t look common to me! Make sure this form gets filled out correctly. Common NameThis is the domain name of the Web site. In my case I’m creating a wildcard certificate so I’m using the * prefix. If you’re purchasing a certificate for a specific domain use www.west-wind.com or store.west-wind.com for example. Make sure this matches the EXACT domain you’re trying to use secure access on because that’s all the certificate is going to work on unless you get a wildcard certificate. Organization Is the name of your company or organization. Depending on the kind of certificate you purchase this name will show up on your certificate. Most low end SSL certificates (ie. those that cost under $100 for single domains) don’t list the organization, the higher signature certificates that also require extensive validation by the cert authority do. Regardless you should make sure this matches the right company/organization. Organizational Unit This can be anything. Not really sure what this is for, but traditionally I’ve always set this to Web because – well this is a Web thing after all right? I’ve never seen this used anywhere that I can tell other than to internally reference the cert. State and CountryPretty obvious. Should reflect the location of the business/organization/person or site.   Next you have to configure the bit size used for the certificate: The default on this dialog is 1024, but I’ve found that most providers these days request a minimum bit length of 2048, as did my DNSimple provider. Again check with the provider when you submit to make sure. Bit length mismatches can cause problems if you use a size that isn’t supported by the provider. I had that happen last year when I submitted my CSR and it got rejected quite a bit later, when the certs usually are issued within an hour or less. When you’re done here, the certificate is saved to disk as a .txt file and it should look something like this (this is a 2048 bit length CSR):-----BEGIN NEW CERTIFICATE REQUEST----- MIIEVGCCAz0CAQAwdjELMAkGA1UEBhMCVVMxDzANBgNVBAgMBkhhd2FpaTENMAsG A1UEBwwEUGFpYTEfMB0GA1UECgwWV2VzdCBXaW5kIFRlY2hub2xvZ2llczEMMAoG B1UECwwDV2ViMRgwFgYDVQQDDA8qLndlc3Qtd2luZC5jb20wggEiMA0GCSqGSIb3 DQEBAQUAA4IBDwAwggEKAoIBAQDIPWOFMkMVRp2Ftj9w/cCVV4OYYhoZYtl+8lTk oqDwKca0xWHLgioX/9v0rZLS6a82MHqKEBxVXu+cuCmSE4AQtB/1YH9lS4tpc/be OZDvnTotP6l4MCEzzAfROcw4CiIg6X0RMSnl8IATAvv2V5LQM9TDdt9oDdMpX2IY +vVC9RZ7PMHBmR9kwI2i/lrKitzhQKaHgpmKcRlM6iqpALUiX28w5HJaDKK1MDHN 607tyFJLHijuJKx7PdTqZYf50KkC3NupfZ2avVycf18Q13jHWj59tvwEOczoVzRL l4LQivAqbhyiqMpWnrZunIOUZta5aGm+jo7O1knGWJjxuraTAgMBAAGgggGYMBoG CisGAQQBgjcNAgMxDBYKNi4yLjkyMDAuMjA0BgkrBgEEAYI3FRQxJzAlAgEFDAZS QVNYUFMMC1JBU1hQU1xSaWNrDAtJbmV0TWdyLmV4ZTByBgorBgEEAYI3DQICMWQw YgIBAR5aAE0AaQBjAHIAbwBzAG8AZgB0ACAAUgBTAEEAIABTAEMAaABhAG4AbgBl AGwAIABDAHIAeQBwAHQAbwBnAHIAYQBwAGgAaQBjACAAUAByAG8AdgBpAGQAZQBy AwEAMIHPBgkqhkiG9w0BCQ4xgcEwgb4wDgYDVR0PAQH/BAQDAgTwMBMGA1UdJQQM MAoGCCsGAQUFBwMBMHgGCSqGSIb3DQEJDwRrMGkwDgYIKoZIhvcNAwICAgCAMA4G CCqGSIb3DQMEAgIAgDALBglghkgBZQMEASowCwYJYIZIAWUDBAEtMAsGCWCGSAFl AwQBAjALBglghkgBZQMEAQUwBwYFKw4DAgcwCgYIKoZIhvcNAwcwHQYDVR0OBBYE FD/yOsTbXE+GVFCFMmldzQvyloz9MA0GCSqGSIb3DQEBBQUAA4IBAQCK6LlsCuIM 1AU0niB6QZ9v0FTsGFxP1dYvVUnJyY6VEKNiGFiQjZac7UCs0p58yScdXWEFOE8V OsjAYD3xYNc05+ckyD67UHRGEUAVB9RBvbKW23KeR/8kBmEzc8PemD52YOgExxAJ 57xWmAwEHAvbgYzQvhO8AOzH3TGvvHbg5UKM1pYgNmuwZq5DkL/IDoeIJwfk/wrI wghNTuxxIFgbH4YrgLgv4PRvrS/LaTCRBdboaCgzATMczaOb1nd/DVNR+3fCtMhM W0psTAjzRbmXF3nJyAQa7jF/52gkY0RfFX2lG5tJnG+XDsVNvKNvh9Qa5Tlmkm06 ILKCm9ciWCKk -----END NEW CERTIFICATE REQUEST----- You can take that certificate request and submit that to your certificate provider. Since this is base64 encoded you can typically just paste it into a text box on the submission page, or some providers will ask you to upload the CSR as a file. What does a Renewal look like? Note the length of the CSR will vary somewhat with key strength, but compare this to a renewal request that IIS generated from my existing site:-----BEGIN NEW CERTIFICATE REQUEST----- MIIPpwYFKoZIhvcNAQcCoIIPmDCCD5QCAQExCzAJBgUrDgMCGgUAMIIIqAYJKoZI hvcNAQcBoIIImQSCCJUwggiRMIIH+gIBADBdMSEwHwYDVQQLDBhEb21haW4gQ29u dHJvbCBWYWxpFGF0ZWQxHjAcBgNVBAsMFUVzc2VudGlhbFNTTCBXaWxkY2FyZDEY MBYGA1UEAwwPKi53ZXN0LXdpbmQuY29tMIGfMA0GCSqGSIb3DQEBAQUAA4GNADCB iQKBgQCK4OuIOR18Wb8tNMGRZiD1c9X57b332Lj7DhbckFqLs0ys8kVDHrTXSj+T Ye9nmAvfPpZmBtE5p9qRNN79rUYugAdl+qEtE4IJe1bRfxXzcKa1SXa8+TEs3zQa zYSmcR2dDuC8om1eAdeCtt0NnkvANgm1VLwGOor/UHMASaEhCQIDAQABoIIG8jAa BgorBgEEAYI3DQIDMQwWCjYuMi45MjAwLjIwNAYJKwYBBAGCNxUUMScwJQIBBQwG UkFTWFBTDAtSQVNYUFNcUmljawwLSW5ldE1nci5leGUwZgYKKwYBBAGCNw0CAjFY MFYCAQIeTgBNAGkAYwByAG8AcwBvAGYAdAAgAFMAdAByAG8AbgBnACAAQwByAHkA cAB0AG8AZwByAGEAcABoAGkAYwAgAFAAcgBvAHYAaQBkAGUAcgMBADCCAQAGCSqG SIb3DQEJDjGB8jCB7zAOBgNVHQ8BAf8EBAMCBaAwDAYDVR0TAQH/BAIwADA0BgNV 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And it didn’t work. IIS creates a custom CSR that is encoded in a format that no certificate authority I’ve ever used uses. If you want the gory details of what’s in there look at this ServerFault question (thanks to Mika in the comments). In the end it doesn’t matter  though – no certificate authority knows what to do with this CSR. So create a new CSR and skip the renewal. Always! Use the same Server Keep in mind that on IIS at least you should always create your certificate on a single server and then when you receive the final certificate from your provider import it on that server. IIS tracks the CSR it created and requires it in order to import the final certificate properly. So if for some reason you try to install the certificate on another server, it won’t work. I’ve also run into trouble trying to install the same certificate twice – this time around I didn’t give my certificate the proper friendly name and IIS failed to allow me to assign the certificate to any of my Web sites. So I removed the certificate and tried to import again, only to find it failed the second time around. There are other ways to fix this, but in my case I had to have the certificate re-issued to work – not what you want to do. Regardless of what you do though, when you import make sure you do it right the first time by crossing all your t’s and dotting your i's– it’ll save you a lot of grief! You don’t actually have to use the server that the certificate gets installed on to generate the CSR and first install it, but it is generally a good idea to do so just so you can get the certificate installed into the right place right away. If you have access to the server where you need to install the certificate you might as well use it. But you can use another machine to generated the and install the certificate, then export the certificate and move it to another machine as needed. So you can use your Dev machine to create a certificate then export it and install it on a live server. More on installation and back up/export later. Installing the Certificate Once you’ve submitted a CSR request your provider will process the request and eventually issue you a new final certificate that contains another text file with the final key to import into your certificate store. IIS does this by combining the content in your certificate request with the original CSR. If all goes well your new certificate shows up in the certificate list and you’re ready to assign the certificate to your sites. Make sure you use a friendly name that matches domain name of your site. So use *.mysite.com or www.mysite.com or store.mysite.com to ensure IIS recognizes the certificate. I made the mistake of not naming my friendly name this way and found that IIS was unable to link my sites to my wildcard certificate. It needed to have the *. as part of the certificate otherwise the Hostname input field was blanked out. Changing the Friendly Name If you by accidentally used an invalid friendly name you can change it later in the Windows certificate store. Bring up a Run Box Type MMC File | Add/Remove Snap In Add Certificates | Computer Account | Local Computer Drill into Certificates | Personal | Certificates Find your Certificate | Right Click | Properties Edit the Friendly Name | Click OK Backing up your Certificate The first thing you should do once your certificate is successfully installed is to back it up! In case your server crashes or you otherwise lose your configuration this will ensure you have an easy way to recover and reinstall your certificate either on the same server or a different one. If you’re running a server farm or using a wildcard certificate you also need to get the certificate onto other machines and a PFX file import is the easiest way to do this. To back up your certificate select your certificate and choose Export from the context or sidebar menu: The Export Certificate option allows you to export a password protected binary file that you can import in a single step. You can copy the resulting binary PFX file to back up or copy to other machines to install on. Importing the certificate on another machine is as easy as pointing at the PFX file and specifying the password. IIS handles the rest. Assigning a new certificate to your Site Once you have the new certificate installed, all that’s left to do is assign it to your site. In IIS select your Web site and bring up the Site Bindings from the right sidebar. Add a new binding for https, bind it to port 443, specify your hostname and pick the certificate from the pick list. If you’re using a root site make sure to set up your certificate for www.yoursite.com and also for yoursite.com so that both work properly with SSL. Note that you need to explicitly configure each hostname for a certificate if you plan to use SSL. Luckily if you update your SSL certificate in the following year, IIS prompts you and asks whether you like to update all other sites that are using the existing cert to the newer cert. And you’re done. So what’s the Pain? So, all of this is old hat and it doesn’t look all that bad right? So what’s the pain here? Well if you follow the instructions and do everything right, then the process is about as straight forward as you would expect it to be. You create a cert request, you import it and assign it to your sites. That’s the basic steps and to be perfectly fair it works well – if nothing goes wrong. However, renewing tends to be the problem. The first unintuitive issue is that you simply shouldn’t renew but create a new CSR and generate your new certificate from that. Over the years I’ve fallen prey to the belief that Microsoft eventually will fix this so that the renewal creates the same type of CSR as the old cert, but apparently that will just never happen. Booo! The other problem I ran into is that I accidentally misnamed my imported certificate which in turn set off a chain of events that caused my originally issued certificate to become uninstallable. When I received my completed certificate I installed it and it installed just fine, but the friendly name was wrong. As a result IIS refused to assign the certificate to any of my host headered sites. That’s strike number one. Why the heck should the friendly name have any effect on the ability to attach the certificate??? Next I uninstalled the certificate because I figured that would be the easiest way to make sure I get it right. But I found that I could not reinstall my certificate. I kept getting these stop errors: "ASN1 bad tag value met" that would prevent the installation from completion. After searching around for this error and reading countless long messages on forums, I found that this error supposedly does not actually mean the install failed, but the list wouldn’t refresh. Commodo has this to say: Note: There is a known issue in IIS 7 giving the following error: "Cannot find the certificate request associated with this certificate file. A certificate request must be completed on the computer where it was created." You may also receive a message stating "ASN1 bad tag value met". If this is the same server that you generated the CSR on then, in most cases, the certificate is actually installed. Simply cancel the dialog and press "F5" to refresh the list of server certificates. If the new certificate is now in the list, you can continue with the next step. If it is not in the list, you will need to reissue your certificate using a new CSR (see our CSR creation instructions for IIS 7). After creating a new CSR, login to your Comodo account and click the 'replace' button for your certificate. Not sure if this issue is fixed in IIS 8 but that’s an insane bug to have crop up. As it turns out, in my case the refresh didn’t work and the certificate didn’t show up in the IIS list after the reinstall. In fact when looking at the certificate store I could see my certificate was installed in the right place, but the private key is missing which is most likely why IIS is not picking it up. It looks like IIS could not match the final cert to the original CSR generated. But again some sort of message to that affect might be helpful instead of ASN1 bad tag value met. Recovering the Private Key So it turns out my original problem was that I received the published key, but when I imported the private key was missing. There’s a relatively easy way to recover from this. If your certificate doesn’t show up in IIS check in the certificate store for the local machine (see steps above on how to bring this up). If you look at the certificate in Certificates/Personal/Certificates make sure you see the key as shown in the image below: if the key is missing it means that the certificate is missing the private key most likely. To fix a certificate you can do the following: Double click the certificate Go to the Details Tab Copy down the Serial number You can copy the serial number from the area blurred out above. The serial number will be in a format like ?00 a7 9b a1 a4 9d 91 63 57 d6 9f 26 b8 ee 79 b5 cb and you’ll need to strip out the spaces in order to use it in the next step. Next open up an Administrative command prompt and issue the following command: certutil -repairstore my 00a79ba1a49d916357d69f26b8ee79b5cb You should get a confirmation message that the repair worked. If you now go back to the certificate store you should now see the key icon show up on the certificate. Your certificate is fixed. Now go back into IIS Manager and refresh the list of certificates and if all goes well you should see all the certificates that showed in the cert store now: Remember – back up the key first then map to your site… Summary I deal with a lot of customers who run their own IIS servers, and I can’t tell you how often I hear about botched SSL installations. When I posted some of my issues on Twitter yesterday I got a hell storm of “me too” responses. I’m clearly not the only one, who’s run into this especially with renewals. I feel pretty comfortable with IIS configuration and I do a lot of it for support purposes, but the SSL configuration is one that never seems to go seamlessly. This blog post is meant as reminder to myself to read next time I do a renewal. So I can dot my i's and dash my t’s before I get caught in the mess I’m dealing with today. Hopefully some of you find this useful as well.© Rick Strahl, West Wind Technologies, 2005-2014Posted in IIS7  Security   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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  • Deploy ASP.NET Web Applications with Web Deployment Projects

    - by Ben Griswold
    One may quickly build and deploy an ASP.NET web application via the Publish option in Visual Studio.  This option works great for most simple deployment scenarios but it won’t always cut it.  Let’s say you need to automate your deployments. Or you have environment-specific configuration settings. Or you need to execute pre/post build operations when you do your builds.  If so, you should consider using Web Deployment Projects. The Web Deployment Project type doesn’t come out-of-the-box with Visual Studio 2008.  You’ll need to Download Visual Studio® 2008 Web Deployment Projects – RTW and install if you want to follow along with this tutorial. I’ve created a shiny new ASP.NET MVC project.  Web Deployment Projects work with websites, web applications and MVC projects so feel free to go with any web project type you’d like.  Once your web application is in place, it’s time to add the Web Deployment project.  You can hunt and peck around the File > New > New Project… dialogue as long as you’d like, but you aren’t going to find what you need.  Instead, select the web project and then choose the “Add Web Deployment Project…” hiding behind the Build menu option. I prefer to name my projects based on the environment in which I plan to deploy.  In this case, I’ll be rolling to the QA machine. Don’t expect too much to happen at this point.  A seemingly empty project with a funny icon will be added to your solution.  That’s it. I want to take a minute and talk about configuration settings before we continue.  Some of the common settings which might change from environment to environment are appSettings, connectionStrings and mailSettings.  Here’s a look at my updated web.config: <appSettings>   <add key="MvcApplication293.Url" value="http://localhost:50596/" />     </appSettings> <connectionStrings>   <add name="ApplicationServices"        connectionString="data source=.\SQLEXPRESS;Integrated Security=SSPI;AttachDBFilename=|DataDirectory|aspnetdb.mdf;User Instance=true"        providerName="System.Data.SqlClient"/> </connectionStrings>   <system.net>   <mailSettings>     <smtp from="[email protected]">         <network host="server.com" userName="username" password="password" port="587" defaultCredentials="false"/>     </smtp>   </mailSettings> </system.net> I want to update these values prior to deploying to the QA environment.  There are variations to this approach, but I like to maintain environment-specific settings for each of the web.config sections in the Config/[Environment] project folders.  I’ve provided a screenshot of the QA environment settings below. It may be obvious what one should include in each of the three files.  Basically, it is a copy of the associated web.config section with updated setting values.  For example, the AppSettings.config file may include a reference to the QA web url, the DB.config would include the QA database server and login information and the StmpSettings.config would include a QA Stmp server and user information. <?xml version="1.0" encoding="utf-8" ?> <appSettings>   <add key="MvcApplication293.Url" value="http://qa.MvcApplicatinon293.com/" /> </appSettings> AppSettings.config  <?xml version="1.0" encoding="utf-8" ?> <connectionStrings>   <add name="ApplicationServices"        connectionString="server=QAServer;integrated security=SSPI;database=MvcApplication293"        providerName="System.Data.SqlClient"/>   </connectionStrings> Db.config  <?xml version="1.0" encoding="utf-8" ?> <smtp from="[email protected]">     <network host="qaserver.com" userName="qausername" password="qapassword" port="587" defaultCredentials="false"/> </smtp> SmtpSettings.config  I think our web project is ready to deploy.  Now, it’s time to concentrate on the Web Deployment Project itself.  Right-click on the project file and open the Property Pages. The first thing to call out is the Configuration dropdown.  I only deploy a project which is built in Release Mode so I only setup the Web Deployment Project for this mode.  (This is when you change the Configuration selection to “Release.”)  I typically keep the Output Folder default value – .\Release\.  When the application is built, all artifacts will be dropped in the .\Release\ folder relative to the Web Deployment Project root.  The final option may be up for some debate.  I like to roll out updatable websites so I select the “Allow this precompiled site to be updatable” option.  I really do like to follow standard SDLC processes when I release my software but there are those times when you just have to make a hotfix to production and I like to keep this option open if need be.  If you are strongly opposed to this idea, please, by all means, don’t check the box. The next tab is boring.  I don’t like to deploy a crazy number of DLLs so I merge all outputs to a single assembly.  Again, you may have another option and feel free to change this selection if you so wish. If you follow my lead, take care when choosing a single assembly name.  The Assembly Name can not be the same as the website or any other project in your solution otherwise you’ll receive a circular reference build error.  In other words, I can’t name the assembly MvcApplication293 or my output window would start yelling at me. Remember when we called out our QA configuration files?  Click on the Deployment tab and you’ll see how where going to use them.  Notice the Web.config file section replacements value.  All this does is swap called out web.config sections with the content of the Config\QA\* files.  You can reduce or extend this list as you deem fit.  Did you see the “Use external configuration source file” option?  You know how you can point any of your web.config sections to an external file via the configSource attribute?  This option allows you to leverage that technique and instead of replacing the content of the sections, you will replace the configSource attribute value instead. <appSettings configSource="Config\QA\AppSettings.config" /> Go ahead and Apply your changes.  I’d like to take a look at the project file we just updated.  Right-click on the Web Deployment Project and select “Open Project File.” One of the first configuration blocks reflects core Release build settings.  There are a couple of points I’d like to call out here: DebugSymbols=false ensures the compilation debug attribute in your web.config is flipped to false as part of build process.  There’s some crumby (more likely old) documentation which implies you need a ToggleDebugCompilation task to make this happen.  Nope. Just make sure the DebugSymbols is set to false.  EnableUpdateable implies a single dll for the web application rather than a dll for each object and and empty view file. I think updatable applications are cleaner and include the benefit (or risk based on your perspective) that portions of the application can be updated directly on the server.  I called this out earlier but I wanted to reiterate. <PropertyGroup Condition=" '$(Configuration)|$(Platform)' == 'Release|AnyCPU' ">     <DebugSymbols>false</DebugSymbols>     <OutputPath>.\Release</OutputPath>     <EnableUpdateable>true</EnableUpdateable>     <UseMerge>true</UseMerge>     <SingleAssemblyName>MvcApplication293</SingleAssemblyName>     <DeleteAppCodeCompiledFiles>true</DeleteAppCodeCompiledFiles>     <UseWebConfigReplacement>true</UseWebConfigReplacement>     <ValidateWebConfigReplacement>true</ValidateWebConfigReplacement>     <DeleteAppDataFolder>true</DeleteAppDataFolder>   </PropertyGroup> The next section is self-explanatory.  The content merely reflects the replacement value you provided via the Property Pages. <ItemGroup Condition="'$(Configuration)|$(Platform)' == 'Release|AnyCPU'">     <WebConfigReplacementFiles Include="Config\QA\AppSettings.config">       <Section>appSettings</Section>     </WebConfigReplacementFiles>     <WebConfigReplacementFiles Include="Config\QA\Db.config">       <Section>connectionStrings</Section>     </WebConfigReplacementFiles>     <WebConfigReplacementFiles Include="Config\QA\SmtpSettings.config">       <Section>system.net/mailSettings/smtp</Section>     </WebConfigReplacementFiles>   </ItemGroup> You’ll want to extend the ItemGroup section to include the files you wish to exclude from the build.  The sample ExcludeFromBuild nodes exclude all obj, svn, csproj, user, pdb artifacts from the build. Enough though they files aren’t included in your web project, you’ll need to exclude them or they’ll show up along with required deployment artifacts.  <ItemGroup Condition="'$(Configuration)|$(Platform)' == 'Release|AnyCPU'">     <WebConfigReplacementFiles Include="Config\QA\AppSettings.config">       <Section>appSettings</Section>     </WebConfigReplacementFiles>     <WebConfigReplacementFiles Include="Config\QA\Db.config">       <Section>connectionStrings</Section>     </WebConfigReplacementFiles>     <WebConfigReplacementFiles Include="Config\QA\SmtpSettings.config">       <Section>system.net/mailSettings/smtp</Section>     </WebConfigReplacementFiles>     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\obj\**\*.*" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\**\.svn\**\*.*" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\**\.svn\**\*" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\**\*.csproj" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\**\*.user" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\bin\*.pdb" />     <ExcludeFromBuild Include="$(SourceWebPhysicalPath)\Notes.txt" />   </ItemGroup> Pre/post build and Pre/post merge tasks are added to the final code block.  By default, your project file should look like the following – a completely commented out section. <!– To modify your build process, add your task inside one of        the targets below and uncomment it. Other similar extension        points exist, see Microsoft.WebDeployment.targets.   <Target Name="BeforeBuild">   </Target>   <Target Name="BeforeMerge">   </Target>   <Target Name="AfterMerge">   </Target>   <Target Name="AfterBuild">   </Target>   –> Update the section to remove all temporary Config folders and files after the build.  <!– To modify your build process, add your task inside one of        the targets below and uncomment it. Other similar extension        points exist, see Microsoft.WebDeployment.targets.     <Target Name="BeforeMerge">   </Target>   <Target Name="AfterMerge">   </Target>     <Target Name="BeforeBuild">      </Target>       –>   <Target Name="AfterBuild">     <!– WebConfigReplacement requires the Config files. Remove after build. –>     <RemoveDir Directories="$(OutputPath)\Config" />   </Target> That’s it for setup.  Save the project file, flip the solution to Release Mode and build.  If there’s an issue, consult the Output window for details.  If all went well, you will find your deployment artifacts in your Web Deployment Project folder like so. Both the code source and published application will be there. Inside the Release folder you will find your “published files” and you’ll notice the Config folder is no where to be found.  In the Source folder, all project files are found with the exception of the items which were excluded from the build. I’ll wrap up this tutorial by calling out a little Web Deployment pet peeve of mine: there doesn’t appear to be a way to add an existing web deployment project to a solution.  The best I can come up with is create a new web deployment project and then copy and paste the contents of the existing project file into the new project file.  It’s not a big deal but it bugs me. Download the Solution

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  • Creating a dynamic proxy generator with c# – Part 3 – Creating the constructors

    - by SeanMcAlinden
    Creating a dynamic proxy generator with c# – Part 1 – Creating the Assembly builder, Module builder and caching mechanism Creating a dynamic proxy generator with c# – Part 2 – Interceptor Design For the latest code go to http://rapidioc.codeplex.com/ When building our proxy type, the first thing we need to do is build the constructors. There needs to be a corresponding constructor for each constructor on the passed in base type. We also want to create a field to store the interceptors and construct this list within each constructor. So assuming the passed in base type is a User<int, IRepository> class, were looking to generate constructor code like the following:   Default Constructor public User`2_RapidDynamicBaseProxy() {     this.interceptors = new List<IInterceptor<User<int, IRepository>>>();     DefaultInterceptor<User<int, IRepository>> item = new DefaultInterceptor<User<int, IRepository>>();     this.interceptors.Add(item); }     Parameterised Constructor public User`2_RapidDynamicBaseProxy(IRepository repository1) : base(repository1) {     this.interceptors = new List<IInterceptor<User<int, IRepository>>>();     DefaultInterceptor<User<int, IRepository>> item = new DefaultInterceptor<User<int, IRepository>>();     this.interceptors.Add(item); }   As you can see, we first populate a field on the class with a new list of the passed in base type. Construct our DefaultInterceptor class. Add the DefaultInterceptor instance to our interceptor collection. Although this seems like a relatively small task, there is a fair amount of work require to get this going. Instead of going through every line of code – please download the latest from http://rapidioc.codeplex.com/ and debug through. In this post I’m going to concentrate on explaining how it works. TypeBuilder The TypeBuilder class is the main class used to create the type. You instantiate a new TypeBuilder using the assembly module we created in part 1. /// <summary> /// Creates a type builder. /// </summary> /// <typeparam name="TBase">The type of the base class to be proxied.</typeparam> public static TypeBuilder CreateTypeBuilder<TBase>() where TBase : class {     TypeBuilder typeBuilder = DynamicModuleCache.Get.DefineType         (             CreateTypeName<TBase>(),             TypeAttributes.Class | TypeAttributes.Public,             typeof(TBase),             new Type[] { typeof(IProxy) }         );       if (typeof(TBase).IsGenericType)     {         GenericsHelper.MakeGenericType(typeof(TBase), typeBuilder);     }       return typeBuilder; }   private static string CreateTypeName<TBase>() where TBase : class {     return string.Format("{0}_RapidDynamicBaseProxy", typeof(TBase).Name); } As you can see, I’ve create a new public class derived from TBase which also implements my IProxy interface, this is used later for adding interceptors. If the base type is generic, the following GenericsHelper.MakeGenericType method is called. GenericsHelper using System; using System.Reflection.Emit; namespace Rapid.DynamicProxy.Types.Helpers {     /// <summary>     /// Helper class for generic types and methods.     /// </summary>     internal static class GenericsHelper     {         /// <summary>         /// Makes the typeBuilder a generic.         /// </summary>         /// <param name="concrete">The concrete.</param>         /// <param name="typeBuilder">The type builder.</param>         public static void MakeGenericType(Type baseType, TypeBuilder typeBuilder)         {             Type[] genericArguments = baseType.GetGenericArguments();               string[] genericArgumentNames = GetArgumentNames(genericArguments);               GenericTypeParameterBuilder[] genericTypeParameterBuilder                 = typeBuilder.DefineGenericParameters(genericArgumentNames);               typeBuilder.MakeGenericType(genericTypeParameterBuilder);         }           /// <summary>         /// Gets the argument names from an array of generic argument types.         /// </summary>         /// <param name="genericArguments">The generic arguments.</param>         public static string[] GetArgumentNames(Type[] genericArguments)         {             string[] genericArgumentNames = new string[genericArguments.Length];               for (int i = 0; i < genericArguments.Length; i++)             {                 genericArgumentNames[i] = genericArguments[i].Name;             }               return genericArgumentNames;         }     } }       As you can see, I’m getting all of the generic argument types and names, creating a GenericTypeParameterBuilder and then using the typeBuilder to make the new type generic. InterceptorsField The interceptors field will store a List<IInterceptor<TBase>>. Fields are simple made using the FieldBuilder class. The following code demonstrates how to create the interceptor field. FieldBuilder interceptorsField = typeBuilder.DefineField(     "interceptors",     typeof(System.Collections.Generic.List<>).MakeGenericType(typeof(IInterceptor<TBase>)),       FieldAttributes.Private     ); The field will now exist with the new Type although it currently has no data – we’ll deal with this in the constructor. Add method for interceptorsField To enable us to add to the interceptorsField list, we are going to utilise the Add method that already exists within the System.Collections.Generic.List class. We still however have to create the methodInfo necessary to call the add method. This can be done similar to the following: Add Interceptor Field MethodInfo addInterceptor = typeof(List<>)     .MakeGenericType(new Type[] { typeof(IInterceptor<>).MakeGenericType(typeof(TBase)) })     .GetMethod     (        "Add",        BindingFlags.Instance | BindingFlags.Public | BindingFlags.NonPublic,        null,        new Type[] { typeof(IInterceptor<>).MakeGenericType(typeof(TBase)) },        null     ); So we’ve create a List<IInterceptor<TBase>> type, then using the type created a method info called Add which accepts an IInterceptor<TBase>. Now in our constructor we can use this to call this.interceptors.Add(// interceptor); Building the Constructors This will be the first hard-core part of the proxy building process so I’m going to show the class and then try to explain what everything is doing. For a clear view, download the source from http://rapidioc.codeplex.com/, go to the test project and debug through the constructor building section. Anyway, here it is: DynamicConstructorBuilder using System; using System.Collections.Generic; using System.Reflection; using System.Reflection.Emit; using Rapid.DynamicProxy.Interception; using Rapid.DynamicProxy.Types.Helpers; namespace Rapid.DynamicProxy.Types.Constructors {     /// <summary>     /// Class for creating the proxy constructors.     /// </summary>     internal static class DynamicConstructorBuilder     {         /// <summary>         /// Builds the constructors.         /// </summary>         /// <typeparam name="TBase">The base type.</typeparam>         /// <param name="typeBuilder">The type builder.</param>         /// <param name="interceptorsField">The interceptors field.</param>         public static void BuildConstructors<TBase>             (                 TypeBuilder typeBuilder,                 FieldBuilder interceptorsField,                 MethodInfo addInterceptor             )             where TBase : class         {             ConstructorInfo interceptorsFieldConstructor = CreateInterceptorsFieldConstructor<TBase>();               ConstructorInfo defaultInterceptorConstructor = CreateDefaultInterceptorConstructor<TBase>();               ConstructorInfo[] constructors = typeof(TBase).GetConstructors();               foreach (ConstructorInfo constructorInfo in constructors)             {                 CreateConstructor<TBase>                     (                         typeBuilder,                         interceptorsField,                         interceptorsFieldConstructor,                         defaultInterceptorConstructor,                         addInterceptor,                         constructorInfo                     );             }         }           #region Private Methods           private static void CreateConstructor<TBase>             (                 TypeBuilder typeBuilder,                 FieldBuilder interceptorsField,                 ConstructorInfo interceptorsFieldConstructor,                 ConstructorInfo defaultInterceptorConstructor,                 MethodInfo AddDefaultInterceptor,                 ConstructorInfo constructorInfo             ) where TBase : class         {             Type[] parameterTypes = GetParameterTypes(constructorInfo);               ConstructorBuilder constructorBuilder = CreateConstructorBuilder(typeBuilder, parameterTypes);               ILGenerator cIL = constructorBuilder.GetILGenerator();               LocalBuilder defaultInterceptorMethodVariable =                 cIL.DeclareLocal(typeof(DefaultInterceptor<>).MakeGenericType(typeof(TBase)));               ConstructInterceptorsField(interceptorsField, interceptorsFieldConstructor, cIL);               ConstructDefaultInterceptor(defaultInterceptorConstructor, cIL, defaultInterceptorMethodVariable);               AddDefaultInterceptorToInterceptorsList                 (                     interceptorsField,                     AddDefaultInterceptor,                     cIL,                     defaultInterceptorMethodVariable                 );               CreateConstructor(constructorInfo, parameterTypes, cIL);         }           private static void CreateConstructor(ConstructorInfo constructorInfo, Type[] parameterTypes, ILGenerator cIL)         {             cIL.Emit(OpCodes.Ldarg_0);               if (parameterTypes.Length > 0)             {                 LoadParameterTypes(parameterTypes, cIL);             }               cIL.Emit(OpCodes.Call, constructorInfo);             cIL.Emit(OpCodes.Ret);         }           private static void LoadParameterTypes(Type[] parameterTypes, ILGenerator cIL)         {             for (int i = 1; i <= parameterTypes.Length; i++)             {                 cIL.Emit(OpCodes.Ldarg_S, i);             }         }           private static void AddDefaultInterceptorToInterceptorsList             (                 FieldBuilder interceptorsField,                 MethodInfo AddDefaultInterceptor,                 ILGenerator cIL,                 LocalBuilder defaultInterceptorMethodVariable             )         {             cIL.Emit(OpCodes.Ldarg_0);             cIL.Emit(OpCodes.Ldfld, interceptorsField);             cIL.Emit(OpCodes.Ldloc, defaultInterceptorMethodVariable);             cIL.Emit(OpCodes.Callvirt, AddDefaultInterceptor);         }           private static void ConstructDefaultInterceptor             (                 ConstructorInfo defaultInterceptorConstructor,                 ILGenerator cIL,                 LocalBuilder defaultInterceptorMethodVariable             )         {             cIL.Emit(OpCodes.Newobj, defaultInterceptorConstructor);             cIL.Emit(OpCodes.Stloc, defaultInterceptorMethodVariable);         }           private static void ConstructInterceptorsField             (                 FieldBuilder interceptorsField,                 ConstructorInfo interceptorsFieldConstructor,                 ILGenerator cIL             )         {             cIL.Emit(OpCodes.Ldarg_0);             cIL.Emit(OpCodes.Newobj, interceptorsFieldConstructor);             cIL.Emit(OpCodes.Stfld, interceptorsField);         }           private static ConstructorBuilder CreateConstructorBuilder(TypeBuilder typeBuilder, Type[] parameterTypes)         {             return typeBuilder.DefineConstructor                 (                     MethodAttributes.Public | MethodAttributes.SpecialName | MethodAttributes.RTSpecialName                     | MethodAttributes.HideBySig, CallingConventions.Standard, parameterTypes                 );         }           private static Type[] GetParameterTypes(ConstructorInfo constructorInfo)         {             ParameterInfo[] parameterInfoArray = constructorInfo.GetParameters();               Type[] parameterTypes = new Type[parameterInfoArray.Length];               for (int p = 0; p < parameterInfoArray.Length; p++)             {                 parameterTypes[p] = parameterInfoArray[p].ParameterType;             }               return parameterTypes;         }           private static ConstructorInfo CreateInterceptorsFieldConstructor<TBase>() where TBase : class         {             return ConstructorHelper.CreateGenericConstructorInfo                 (                     typeof(List<>),                     new Type[] { typeof(IInterceptor<TBase>) },                     BindingFlags.Instance | BindingFlags.Public | BindingFlags.NonPublic                 );         }           private static ConstructorInfo CreateDefaultInterceptorConstructor<TBase>() where TBase : class         {             return ConstructorHelper.CreateGenericConstructorInfo                 (                     typeof(DefaultInterceptor<>),                     new Type[] { typeof(TBase) },                     BindingFlags.Instance | BindingFlags.Public | BindingFlags.NonPublic                 );         }           #endregion     } } So, the first two tasks within the class should be fairly clear, we are creating a ConstructorInfo for the interceptorField list and a ConstructorInfo for the DefaultConstructor, this is for instantiating them in each contructor. We then using Reflection get an array of all of the constructors in the base class, we then loop through the array and create a corresponding proxy contructor. Hopefully, the code is fairly easy to follow other than some new types and the dreaded Opcodes. ConstructorBuilder This class defines a new constructor on the type. ILGenerator The ILGenerator allows the use of Reflection.Emit to create the method body. LocalBuilder The local builder allows the storage of data in local variables within a method, in this case it’s the constructed DefaultInterceptor. Constructing the interceptors field The first bit of IL you’ll come across as you follow through the code is the following private method used for constructing the field list of interceptors. private static void ConstructInterceptorsField             (                 FieldBuilder interceptorsField,                 ConstructorInfo interceptorsFieldConstructor,                 ILGenerator cIL             )         {             cIL.Emit(OpCodes.Ldarg_0);             cIL.Emit(OpCodes.Newobj, interceptorsFieldConstructor);             cIL.Emit(OpCodes.Stfld, interceptorsField);         } The first thing to know about generating code using IL is that you are using a stack, if you want to use something, you need to push it up the stack etc. etc. OpCodes.ldArg_0 This opcode is a really interesting one, basically each method has a hidden first argument of the containing class instance (apart from static classes), constructors are no different. This is the reason you can use syntax like this.myField. So back to the method, as we want to instantiate the List in the interceptorsField, first we need to load the class instance onto the stack, we then load the new object (new List<TBase>) and finally we store it in the interceptorsField. Hopefully, that should follow easily enough in the method. In each constructor you would now have this.interceptors = new List<User<int, IRepository>>(); Constructing and storing the DefaultInterceptor The next bit of code we need to create is the constructed DefaultInterceptor. Firstly, we create a local builder to store the constructed type. Create a local builder LocalBuilder defaultInterceptorMethodVariable =     cIL.DeclareLocal(typeof(DefaultInterceptor<>).MakeGenericType(typeof(TBase))); Once our local builder is ready, we then need to construct the DefaultInterceptor<TBase> and store it in the variable. Connstruct DefaultInterceptor private static void ConstructDefaultInterceptor     (         ConstructorInfo defaultInterceptorConstructor,         ILGenerator cIL,         LocalBuilder defaultInterceptorMethodVariable     ) {     cIL.Emit(OpCodes.Newobj, defaultInterceptorConstructor);     cIL.Emit(OpCodes.Stloc, defaultInterceptorMethodVariable); } As you can see, using the ConstructorInfo named defaultInterceptorConstructor, we load the new object onto the stack. Then using the store local opcode (OpCodes.Stloc), we store the new object in the local builder named defaultInterceptorMethodVariable. Add the constructed DefaultInterceptor to the interceptors field collection Using the add method created earlier in this post, we are going to add the new DefaultInterceptor object to the interceptors field collection. Add Default Interceptor private static void AddDefaultInterceptorToInterceptorsList     (         FieldBuilder interceptorsField,         MethodInfo AddDefaultInterceptor,         ILGenerator cIL,         LocalBuilder defaultInterceptorMethodVariable     ) {     cIL.Emit(OpCodes.Ldarg_0);     cIL.Emit(OpCodes.Ldfld, interceptorsField);     cIL.Emit(OpCodes.Ldloc, defaultInterceptorMethodVariable);     cIL.Emit(OpCodes.Callvirt, AddDefaultInterceptor); } So, here’s whats going on. The class instance is first loaded onto the stack using the load argument at index 0 opcode (OpCodes.Ldarg_0) (remember the first arg is the hidden class instance). The interceptorsField is then loaded onto the stack using the load field opcode (OpCodes.Ldfld). We then load the DefaultInterceptor object we stored locally using the load local opcode (OpCodes.Ldloc). Then finally we call the AddDefaultInterceptor method using the call virtual opcode (Opcodes.Callvirt). Completing the constructor The last thing we need to do is complete the constructor. Complete the constructor private static void CreateConstructor(ConstructorInfo constructorInfo, Type[] parameterTypes, ILGenerator cIL)         {             cIL.Emit(OpCodes.Ldarg_0);               if (parameterTypes.Length > 0)             {                 LoadParameterTypes(parameterTypes, cIL);             }               cIL.Emit(OpCodes.Call, constructorInfo);             cIL.Emit(OpCodes.Ret);         }           private static void LoadParameterTypes(Type[] parameterTypes, ILGenerator cIL)         {             for (int i = 1; i <= parameterTypes.Length; i++)             {                 cIL.Emit(OpCodes.Ldarg_S, i);             }         } So, the first thing we do again is load the class instance using the load argument at index 0 opcode (OpCodes.Ldarg_0). We then load each parameter using OpCode.Ldarg_S, this opcode allows us to specify an index position for each argument. We then setup calling the base constructor using OpCodes.Call and the base constructors ConstructorInfo. Finally, all methods are required to return, even when they have a void return. As there are no values on the stack after the OpCodes.Call line, we can safely call the OpCode.Ret to give the constructor a void return. If there was a value, we would have to pop the value of the stack before calling return otherwise, the method would try and return a value. Conclusion This was a slightly hardcore post but hopefully it hasn’t been too hard to follow. The main thing is that a number of the really useful opcodes have been used and now the dynamic proxy is capable of being constructed. If you download the code and debug through the tests at http://rapidioc.codeplex.com/, you’ll be able to create proxies at this point, they cannon do anything in terms of interception but you can happily run the tests, call base methods and properties and also take a look at the created assembly in Reflector. Hope this is useful. The next post should be up soon, it will be covering creating the private methods for calling the base class methods and properties. Kind Regards, Sean.

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  • Improving Partitioned Table Join Performance

    - by Paul White
    The query optimizer does not always choose an optimal strategy when joining partitioned tables. This post looks at an example, showing how a manual rewrite of the query can almost double performance, while reducing the memory grant to almost nothing. Test Data The two tables in this example use a common partitioning partition scheme. The partition function uses 41 equal-size partitions: CREATE PARTITION FUNCTION PFT (integer) AS RANGE RIGHT FOR VALUES ( 125000, 250000, 375000, 500000, 625000, 750000, 875000, 1000000, 1125000, 1250000, 1375000, 1500000, 1625000, 1750000, 1875000, 2000000, 2125000, 2250000, 2375000, 2500000, 2625000, 2750000, 2875000, 3000000, 3125000, 3250000, 3375000, 3500000, 3625000, 3750000, 3875000, 4000000, 4125000, 4250000, 4375000, 4500000, 4625000, 4750000, 4875000, 5000000 ); GO CREATE PARTITION SCHEME PST AS PARTITION PFT ALL TO ([PRIMARY]); There two tables are: CREATE TABLE dbo.T1 ( TID integer NOT NULL IDENTITY(0,1), Column1 integer NOT NULL, Padding binary(100) NOT NULL DEFAULT 0x,   CONSTRAINT PK_T1 PRIMARY KEY CLUSTERED (TID) ON PST (TID) );   CREATE TABLE dbo.T2 ( TID integer NOT NULL, Column1 integer NOT NULL, Padding binary(100) NOT NULL DEFAULT 0x,   CONSTRAINT PK_T2 PRIMARY KEY CLUSTERED (TID, Column1) ON PST (TID) ); The next script loads 5 million rows into T1 with a pseudo-random value between 1 and 5 for Column1. The table is partitioned on the IDENTITY column TID: INSERT dbo.T1 WITH (TABLOCKX) (Column1) SELECT (ABS(CHECKSUM(NEWID())) % 5) + 1 FROM dbo.Numbers AS N WHERE n BETWEEN 1 AND 5000000; In case you don’t already have an auxiliary table of numbers lying around, here’s a script to create one with 10 million rows: CREATE TABLE dbo.Numbers (n bigint PRIMARY KEY);   WITH L0 AS(SELECT 1 AS c UNION ALL SELECT 1), L1 AS(SELECT 1 AS c FROM L0 AS A CROSS JOIN L0 AS B), L2 AS(SELECT 1 AS c FROM L1 AS A CROSS JOIN L1 AS B), L3 AS(SELECT 1 AS c FROM L2 AS A CROSS JOIN L2 AS B), L4 AS(SELECT 1 AS c FROM L3 AS A CROSS JOIN L3 AS B), L5 AS(SELECT 1 AS c FROM L4 AS A CROSS JOIN L4 AS B), Nums AS(SELECT ROW_NUMBER() OVER (ORDER BY (SELECT NULL)) AS n FROM L5) INSERT dbo.Numbers WITH (TABLOCKX) SELECT TOP (10000000) n FROM Nums ORDER BY n OPTION (MAXDOP 1); Table T1 contains data like this: Next we load data into table T2. The relationship between the two tables is that table 2 contains ‘n’ rows for each row in table 1, where ‘n’ is determined by the value in Column1 of table T1. There is nothing particularly special about the data or distribution, by the way. INSERT dbo.T2 WITH (TABLOCKX) (TID, Column1) SELECT T.TID, N.n FROM dbo.T1 AS T JOIN dbo.Numbers AS N ON N.n >= 1 AND N.n <= T.Column1; Table T2 ends up containing about 15 million rows: The primary key for table T2 is a combination of TID and Column1. The data is partitioned according to the value in column TID alone. Partition Distribution The following query shows the number of rows in each partition of table T1: SELECT PartitionID = CA1.P, NumRows = COUNT_BIG(*) FROM dbo.T1 AS T CROSS APPLY (VALUES ($PARTITION.PFT(TID))) AS CA1 (P) GROUP BY CA1.P ORDER BY CA1.P; There are 40 partitions containing 125,000 rows (40 * 125k = 5m rows). The rightmost partition remains empty. The next query shows the distribution for table 2: SELECT PartitionID = CA1.P, NumRows = COUNT_BIG(*) FROM dbo.T2 AS T CROSS APPLY (VALUES ($PARTITION.PFT(TID))) AS CA1 (P) GROUP BY CA1.P ORDER BY CA1.P; There are roughly 375,000 rows in each partition (the rightmost partition is also empty): Ok, that’s the test data done. Test Query and Execution Plan The task is to count the rows resulting from joining tables 1 and 2 on the TID column: SET STATISTICS IO ON; DECLARE @s datetime2 = SYSUTCDATETIME();   SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID;   SELECT DATEDIFF(Millisecond, @s, SYSUTCDATETIME()); SET STATISTICS IO OFF; The optimizer chooses a plan using parallel hash join, and partial aggregation: The Plan Explorer plan tree view shows accurate cardinality estimates and an even distribution of rows across threads (click to enlarge the image): With a warm data cache, the STATISTICS IO output shows that no physical I/O was needed, and all 41 partitions were touched: Running the query without actual execution plan or STATISTICS IO information for maximum performance, the query returns in around 2600ms. Execution Plan Analysis The first step toward improving on the execution plan produced by the query optimizer is to understand how it works, at least in outline. The two parallel Clustered Index Scans use multiple threads to read rows from tables T1 and T2. Parallel scan uses a demand-based scheme where threads are given page(s) to scan from the table as needed. This arrangement has certain important advantages, but does result in an unpredictable distribution of rows amongst threads. The point is that multiple threads cooperate to scan the whole table, but it is impossible to predict which rows end up on which threads. For correct results from the parallel hash join, the execution plan has to ensure that rows from T1 and T2 that might join are processed on the same thread. For example, if a row from T1 with join key value ‘1234’ is placed in thread 5’s hash table, the execution plan must guarantee that any rows from T2 that also have join key value ‘1234’ probe thread 5’s hash table for matches. The way this guarantee is enforced in this parallel hash join plan is by repartitioning rows to threads after each parallel scan. The two repartitioning exchanges route rows to threads using a hash function over the hash join keys. The two repartitioning exchanges use the same hash function so rows from T1 and T2 with the same join key must end up on the same hash join thread. Expensive Exchanges This business of repartitioning rows between threads can be very expensive, especially if a large number of rows is involved. The execution plan selected by the optimizer moves 5 million rows through one repartitioning exchange and around 15 million across the other. As a first step toward removing these exchanges, consider the execution plan selected by the optimizer if we join just one partition from each table, disallowing parallelism: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = 1 AND $PARTITION.PFT(T2.TID) = 1 OPTION (MAXDOP 1); The optimizer has chosen a (one-to-many) merge join instead of a hash join. The single-partition query completes in around 100ms. If everything scaled linearly, we would expect that extending this strategy to all 40 populated partitions would result in an execution time around 4000ms. Using parallelism could reduce that further, perhaps to be competitive with the parallel hash join chosen by the optimizer. This raises a question. If the most efficient way to join one partition from each of the tables is to use a merge join, why does the optimizer not choose a merge join for the full query? Forcing a Merge Join Let’s force the optimizer to use a merge join on the test query using a hint: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (MERGE JOIN); This is the execution plan selected by the optimizer: This plan results in the same number of logical reads reported previously, but instead of 2600ms the query takes 5000ms. The natural explanation for this drop in performance is that the merge join plan is only using a single thread, whereas the parallel hash join plan could use multiple threads. Parallel Merge Join We can get a parallel merge join plan using the same query hint as before, and adding trace flag 8649: SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (MERGE JOIN, QUERYTRACEON 8649); The execution plan is: This looks promising. It uses a similar strategy to distribute work across threads as seen for the parallel hash join. In practice though, performance is disappointing. On a typical run, the parallel merge plan runs for around 8400ms; slower than the single-threaded merge join plan (5000ms) and much worse than the 2600ms for the parallel hash join. We seem to be going backwards! The logical reads for the parallel merge are still exactly the same as before, with no physical IOs. The cardinality estimates and thread distribution are also still very good (click to enlarge): A big clue to the reason for the poor performance is shown in the wait statistics (captured by Plan Explorer Pro): CXPACKET waits require careful interpretation, and are most often benign, but in this case excessive waiting occurs at the repartitioning exchanges. Unlike the parallel hash join, the repartitioning exchanges in this plan are order-preserving ‘merging’ exchanges (because merge join requires ordered inputs): Parallelism works best when threads can just grab any available unit of work and get on with processing it. Preserving order introduces inter-thread dependencies that can easily lead to significant waits occurring. In extreme cases, these dependencies can result in an intra-query deadlock, though the details of that will have to wait for another time to explore in detail. The potential for waits and deadlocks leads the query optimizer to cost parallel merge join relatively highly, especially as the degree of parallelism (DOP) increases. This high costing resulted in the optimizer choosing a serial merge join rather than parallel in this case. The test results certainly confirm its reasoning. Collocated Joins In SQL Server 2008 and later, the optimizer has another available strategy when joining tables that share a common partition scheme. This strategy is a collocated join, also known as as a per-partition join. It can be applied in both serial and parallel execution plans, though it is limited to 2-way joins in the current optimizer. Whether the optimizer chooses a collocated join or not depends on cost estimation. The primary benefits of a collocated join are that it eliminates an exchange and requires less memory, as we will see next. Costing and Plan Selection The query optimizer did consider a collocated join for our original query, but it was rejected on cost grounds. The parallel hash join with repartitioning exchanges appeared to be a cheaper option. There is no query hint to force a collocated join, so we have to mess with the costing framework to produce one for our test query. Pretending that IOs cost 50 times more than usual is enough to convince the optimizer to use collocated join with our test query: -- Pretend IOs are 50x cost temporarily DBCC SETIOWEIGHT(50);   -- Co-located hash join SELECT COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID OPTION (RECOMPILE);   -- Reset IO costing DBCC SETIOWEIGHT(1); Collocated Join Plan The estimated execution plan for the collocated join is: The Constant Scan contains one row for each partition of the shared partitioning scheme, from 1 to 41. The hash repartitioning exchanges seen previously are replaced by a single Distribute Streams exchange using Demand partitioning. Demand partitioning means that the next partition id is given to the next parallel thread that asks for one. My test machine has eight logical processors, and all are available for SQL Server to use. As a result, there are eight threads in the single parallel branch in this plan, each processing one partition from each table at a time. Once a thread finishes processing a partition, it grabs a new partition number from the Distribute Streams exchange…and so on until all partitions have been processed. It is important to understand that the parallel scans in this plan are different from the parallel hash join plan. Although the scans have the same parallelism icon, tables T1 and T2 are not being co-operatively scanned by multiple threads in the same way. Each thread reads a single partition of T1 and performs a hash match join with the same partition from table T2. The properties of the two Clustered Index Scans show a Seek Predicate (unusual for a scan!) limiting the rows to a single partition: The crucial point is that the join between T1 and T2 is on TID, and TID is the partitioning column for both tables. A thread that processes partition ‘n’ is guaranteed to see all rows that can possibly join on TID for that partition. In addition, no other thread will see rows from that partition, so this removes the need for repartitioning exchanges. CPU and Memory Efficiency Improvements The collocated join has removed two expensive repartitioning exchanges and added a single exchange processing 41 rows (one for each partition id). Remember, the parallel hash join plan exchanges had to process 5 million and 15 million rows. The amount of processor time spent on exchanges will be much lower in the collocated join plan. In addition, the collocated join plan has a maximum of 8 threads processing single partitions at any one time. The 41 partitions will all be processed eventually, but a new partition is not started until a thread asks for it. Threads can reuse hash table memory for the new partition. The parallel hash join plan also had 8 hash tables, but with all 5,000,000 build rows loaded at the same time. The collocated plan needs memory for only 8 * 125,000 = 1,000,000 rows at any one time. Collocated Hash Join Performance The collated join plan has disappointing performance in this case. The query runs for around 25,300ms despite the same IO statistics as usual. This is much the worst result so far, so what went wrong? It turns out that cardinality estimation for the single partition scans of table T1 is slightly low. The properties of the Clustered Index Scan of T1 (graphic immediately above) show the estimation was for 121,951 rows. This is a small shortfall compared with the 125,000 rows actually encountered, but it was enough to cause the hash join to spill to physical tempdb: A level 1 spill doesn’t sound too bad, until you realize that the spill to tempdb probably occurs for each of the 41 partitions. As a side note, the cardinality estimation error is a little surprising because the system tables accurately show there are 125,000 rows in every partition of T1. Unfortunately, the optimizer uses regular column and index statistics to derive cardinality estimates here rather than system table information (e.g. sys.partitions). Collocated Merge Join We will never know how well the collocated parallel hash join plan might have worked without the cardinality estimation error (and the resulting 41 spills to tempdb) but we do know: Merge join does not require a memory grant; and Merge join was the optimizer’s preferred join option for a single partition join Putting this all together, what we would really like to see is the same collocated join strategy, but using merge join instead of hash join. Unfortunately, the current query optimizer cannot produce a collocated merge join; it only knows how to do collocated hash join. So where does this leave us? CROSS APPLY sys.partitions We can try to write our own collocated join query. We can use sys.partitions to find the partition numbers, and CROSS APPLY to get a count per partition, with a final step to sum the partial counts. The following query implements this idea: SELECT row_count = SUM(Subtotals.cnt) FROM ( -- Partition numbers SELECT p.partition_number FROM sys.partitions AS p WHERE p.[object_id] = OBJECT_ID(N'T1', N'U') AND p.index_id = 1 ) AS P CROSS APPLY ( -- Count per collocated join SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals; The estimated plan is: The cardinality estimates aren’t all that good here, especially the estimate for the scan of the system table underlying the sys.partitions view. Nevertheless, the plan shape is heading toward where we would like to be. Each partition number from the system table results in a per-partition scan of T1 and T2, a one-to-many Merge Join, and a Stream Aggregate to compute the partial counts. The final Stream Aggregate just sums the partial counts. Execution time for this query is around 3,500ms, with the same IO statistics as always. This compares favourably with 5,000ms for the serial plan produced by the optimizer with the OPTION (MERGE JOIN) hint. This is another case of the sum of the parts being less than the whole – summing 41 partial counts from 41 single-partition merge joins is faster than a single merge join and count over all partitions. Even so, this single-threaded collocated merge join is not as quick as the original parallel hash join plan, which executed in 2,600ms. On the positive side, our collocated merge join uses only one logical processor and requires no memory grant. The parallel hash join plan used 16 threads and reserved 569 MB of memory:   Using a Temporary Table Our collocated merge join plan should benefit from parallelism. The reason parallelism is not being used is that the query references a system table. We can work around that by writing the partition numbers to a temporary table (or table variable): SET STATISTICS IO ON; DECLARE @s datetime2 = SYSUTCDATETIME();   CREATE TABLE #P ( partition_number integer PRIMARY KEY);   INSERT #P (partition_number) SELECT p.partition_number FROM sys.partitions AS p WHERE p.[object_id] = OBJECT_ID(N'T1', N'U') AND p.index_id = 1;   SELECT row_count = SUM(Subtotals.cnt) FROM #P AS p CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals;   DROP TABLE #P;   SELECT DATEDIFF(Millisecond, @s, SYSUTCDATETIME()); SET STATISTICS IO OFF; Using the temporary table adds a few logical reads, but the overall execution time is still around 3500ms, indistinguishable from the same query without the temporary table. The problem is that the query optimizer still doesn’t choose a parallel plan for this query, though the removal of the system table reference means that it could if it chose to: In fact the optimizer did enter the parallel plan phase of query optimization (running search 1 for a second time): Unfortunately, the parallel plan found seemed to be more expensive than the serial plan. This is a crazy result, caused by the optimizer’s cost model not reducing operator CPU costs on the inner side of a nested loops join. Don’t get me started on that, we’ll be here all night. In this plan, everything expensive happens on the inner side of a nested loops join. Without a CPU cost reduction to compensate for the added cost of exchange operators, candidate parallel plans always look more expensive to the optimizer than the equivalent serial plan. Parallel Collocated Merge Join We can produce the desired parallel plan using trace flag 8649 again: SELECT row_count = SUM(Subtotals.cnt) FROM #P AS p CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals OPTION (QUERYTRACEON 8649); The actual execution plan is: One difference between this plan and the collocated hash join plan is that a Repartition Streams exchange operator is used instead of Distribute Streams. The effect is similar, though not quite identical. The Repartition uses round-robin partitioning, meaning the next partition id is pushed to the next thread in sequence. The Distribute Streams exchange seen earlier used Demand partitioning, meaning the next partition id is pulled across the exchange by the next thread that is ready for more work. There are subtle performance implications for each partitioning option, but going into that would again take us too far off the main point of this post. Performance The important thing is the performance of this parallel collocated merge join – just 1350ms on a typical run. The list below shows all the alternatives from this post (all timings include creation, population, and deletion of the temporary table where appropriate) from quickest to slowest: Collocated parallel merge join: 1350ms Parallel hash join: 2600ms Collocated serial merge join: 3500ms Serial merge join: 5000ms Parallel merge join: 8400ms Collated parallel hash join: 25,300ms (hash spill per partition) The parallel collocated merge join requires no memory grant (aside from a paltry 1.2MB used for exchange buffers). This plan uses 16 threads at DOP 8; but 8 of those are (rather pointlessly) allocated to the parallel scan of the temporary table. These are minor concerns, but it turns out there is a way to address them if it bothers you. Parallel Collocated Merge Join with Demand Partitioning This final tweak replaces the temporary table with a hard-coded list of partition ids (dynamic SQL could be used to generate this query from sys.partitions): SELECT row_count = SUM(Subtotals.cnt) FROM ( VALUES (1),(2),(3),(4),(5),(6),(7),(8),(9),(10), (11),(12),(13),(14),(15),(16),(17),(18),(19),(20), (21),(22),(23),(24),(25),(26),(27),(28),(29),(30), (31),(32),(33),(34),(35),(36),(37),(38),(39),(40),(41) ) AS P (partition_number) CROSS APPLY ( SELECT cnt = COUNT_BIG(*) FROM dbo.T1 AS T1 JOIN dbo.T2 AS T2 ON T2.TID = T1.TID WHERE $PARTITION.PFT(T1.TID) = p.partition_number AND $PARTITION.PFT(T2.TID) = p.partition_number ) AS SubTotals OPTION (QUERYTRACEON 8649); The actual execution plan is: The parallel collocated hash join plan is reproduced below for comparison: The manual rewrite has another advantage that has not been mentioned so far: the partial counts (per partition) can be computed earlier than the partial counts (per thread) in the optimizer’s collocated join plan. The earlier aggregation is performed by the extra Stream Aggregate under the nested loops join. The performance of the parallel collocated merge join is unchanged at around 1350ms. Final Words It is a shame that the current query optimizer does not consider a collocated merge join (Connect item closed as Won’t Fix). The example used in this post showed an improvement in execution time from 2600ms to 1350ms using a modestly-sized data set and limited parallelism. In addition, the memory requirement for the query was almost completely eliminated  – down from 569MB to 1.2MB. The problem with the parallel hash join selected by the optimizer is that it attempts to process the full data set all at once (albeit using eight threads). It requires a large memory grant to hold all 5 million rows from table T1 across the eight hash tables, and does not take advantage of the divide-and-conquer opportunity offered by the common partitioning. The great thing about the collocated join strategies is that each parallel thread works on a single partition from both tables, reading rows, performing the join, and computing a per-partition subtotal, before moving on to a new partition. From a thread’s point of view… If you have trouble visualizing what is happening from just looking at the parallel collocated merge join execution plan, let’s look at it again, but from the point of view of just one thread operating between the two Parallelism (exchange) operators. Our thread picks up a single partition id from the Distribute Streams exchange, and starts a merge join using ordered rows from partition 1 of table T1 and partition 1 of table T2. By definition, this is all happening on a single thread. As rows join, they are added to a (per-partition) count in the Stream Aggregate immediately above the Merge Join. Eventually, either T1 (partition 1) or T2 (partition 1) runs out of rows and the merge join stops. The per-partition count from the aggregate passes on through the Nested Loops join to another Stream Aggregate, which is maintaining a per-thread subtotal. Our same thread now picks up a new partition id from the exchange (say it gets id 9 this time). The count in the per-partition aggregate is reset to zero, and the processing of partition 9 of both tables proceeds just as it did for partition 1, and on the same thread. Each thread picks up a single partition id and processes all the data for that partition, completely independently from other threads working on other partitions. One thread might eventually process partitions (1, 9, 17, 25, 33, 41) while another is concurrently processing partitions (2, 10, 18, 26, 34) and so on for the other six threads at DOP 8. The point is that all 8 threads can execute independently and concurrently, continuing to process new partitions until the wider job (of which the thread has no knowledge!) is done. This divide-and-conquer technique can be much more efficient than simply splitting the entire workload across eight threads all at once. Related Reading Understanding and Using Parallelism in SQL Server Parallel Execution Plans Suck © 2013 Paul White – All Rights Reserved Twitter: @SQL_Kiwi

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  • Windows Azure: Backup Services Release, Hyper-V Recovery Manager, VM Enhancements, Enhanced Enterprise Management Support

    - by ScottGu
    This morning we released a huge set of updates to Windows Azure.  These new capabilities include: Backup Services: General Availability of Windows Azure Backup Services Hyper-V Recovery Manager: Public preview of Windows Azure Hyper-V Recovery Manager Virtual Machines: Delete Attached Disks, Availability Set Warnings, SQL AlwaysOn Configuration Active Directory: Securely manage hundreds of SaaS applications Enterprise Management: Use Active Directory to Better Manage Windows Azure Windows Azure SDK 2.2: A massive update of our SDK + Visual Studio tooling support All of these improvements are now available to use immediately.  Below are more details about them. Backup Service: General Availability Release of Windows Azure Backup Today we are releasing Windows Azure Backup Service as a general availability service.  This release is now live in production, backed by an enterprise SLA, supported by Microsoft Support, and is ready to use for production scenarios. Windows Azure Backup is a cloud based backup solution for Windows Server which allows files and folders to be backed up and recovered from the cloud, and provides off-site protection against data loss. The service provides IT administrators and developers with the option to back up and protect critical data in an easily recoverable way from any location with no upfront hardware cost. Windows Azure Backup is built on the Windows Azure platform and uses Windows Azure blob storage for storing customer data. Windows Server uses the downloadable Windows Azure Backup Agent to transfer file and folder data securely and efficiently to the Windows Azure Backup Service. Along with providing cloud backup for Windows Server, Windows Azure Backup Service also provides capability to backup data from System Center Data Protection Manager and Windows Server Essentials, to the cloud. All data is encrypted onsite before it is sent to the cloud, and customers retain and manage the encryption key (meaning the data is stored entirely secured and can’t be decrypted by anyone but yourself). Getting Started To get started with the Windows Azure Backup Service, create a new Backup Vault within the Windows Azure Management Portal.  Click New->Data Services->Recovery Services->Backup Vault to do this: Once the backup vault is created you’ll be presented with a simple tutorial that will help guide you on how to register your Windows Servers with it: Once the servers you want to backup are registered, you can use the appropriate local management interface (such as the Microsoft Management Console snap-in, System Center Data Protection Manager Console, or Windows Server Essentials Dashboard) to configure the scheduled backups and to optionally initiate recoveries. You can follow these tutorials to learn more about how to do this: Tutorial: Schedule Backups Using the Windows Azure Backup Agent This tutorial helps you with setting up a backup schedule for your registered Windows Servers. Additionally, it also explains how to use Windows PowerShell cmdlets to set up a custom backup schedule. Tutorial: Recover Files and Folders Using the Windows Azure Backup Agent This tutorial helps you with recovering data from a backup. Additionally, it also explains how to use Windows PowerShell cmdlets to do the same tasks. Below are some of the key benefits the Windows Azure Backup Service provides: Simple configuration and management. Windows Azure Backup Service integrates with the familiar Windows Server Backup utility in Windows Server, the Data Protection Manager component in System Center and Windows Server Essentials, in order to provide a seamless backup and recovery experience to a local disk, or to the cloud. Block level incremental backups. The Windows Azure Backup Agent performs incremental backups by tracking file and block level changes and only transferring the changed blocks, hence reducing the storage and bandwidth utilization. Different point-in-time versions of the backups use storage efficiently by only storing the changes blocks between these versions. Data compression, encryption and throttling. The Windows Azure Backup Agent ensures that data is compressed and encrypted on the server before being sent to the Windows Azure Backup Service over the network. As a result, the Windows Azure Backup Service only stores encrypted data in the cloud storage. The encryption key is not available to the Windows Azure Backup Service, and as a result the data is never decrypted in the service. Also, users can setup throttling and configure how the Windows Azure Backup service utilizes the network bandwidth when backing up or restoring information. Data integrity is verified in the cloud. In addition to the secure backups, the backed up data is also automatically checked for integrity once the backup is done. As a result, any corruptions which may arise due to data transfer can be easily identified and are fixed automatically. Configurable retention policies for storing data in the cloud. The Windows Azure Backup Service accepts and implements retention policies to recycle backups that exceed the desired retention range, thereby meeting business policies and managing backup costs. Hyper-V Recovery Manager: Now Available in Public Preview I’m excited to also announce the public preview of a new Windows Azure Service – the Windows Azure Hyper-V Recovery Manager (HRM). Windows Azure Hyper-V Recovery Manager helps protect your business critical services by coordinating the replication and recovery of System Center Virtual Machine Manager 2012 SP1 and System Center Virtual Machine Manager 2012 R2 private clouds at a secondary location. With automated protection, asynchronous ongoing replication, and orderly recovery, the Hyper-V Recovery Manager service can help you implement Disaster Recovery and restore important services accurately, consistently, and with minimal downtime. Application data in an Hyper-V Recovery Manager scenarios always travels on your on-premise replication channel. Only metadata (such as names of logical clouds, virtual machines, networks etc.) that is needed for orchestration is sent to Azure. All traffic sent to/from Azure is encrypted. You can begin using Windows Azure Hyper-V Recovery today by clicking New->Data Services->Recovery Services->Hyper-V Recovery Manager within the Windows Azure Management Portal.  You can read more about Windows Azure Hyper-V Recovery Manager in Brad Anderson’s 9-part series, Transform the datacenter. To learn more about setting up Hyper-V Recovery Manager follow our detailed step-by-step guide. Virtual Machines: Delete Attached Disks, Availability Set Warnings, SQL AlwaysOn Today’s Windows Azure release includes a number of nice updates to Windows Azure Virtual Machines.  These improvements include: Ability to Delete both VM Instances + Attached Disks in One Operation Prior to today’s release, when you deleted VMs within Windows Azure we would delete the VM instance – but not delete the drives attached to the VM.  You had to manually delete these yourself from the storage account.  With today’s update we’ve added a convenience option that now allows you to either retain or delete the attached disks when you delete the VM:   We’ve also added the ability to delete a cloud service, its deployments, and its role instances with a single action. This can either be a cloud service that has production and staging deployments with web and worker roles, or a cloud service that contains virtual machines.  To do this, simply select the Cloud Service within the Windows Azure Management Portal and click the “Delete” button: Warnings on Availability Sets with Only One Virtual Machine In Them One of the nice features that Windows Azure Virtual Machines supports is the concept of “Availability Sets”.  An “availability set” allows you to define a tier/role (e.g. webfrontends, databaseservers, etc) that you can map Virtual Machines into – and when you do this Windows Azure separates them across fault domains and ensures that at least one of them is always available during servicing operations.  This enables you to deploy applications in a high availability way. One issue we’ve seen some customers run into is where they define an availability set, but then forget to map more than one VM into it (which defeats the purpose of having an availability set).  With today’s release we now display a warning in the Windows Azure Management Portal if you have only one virtual machine deployed in an availability set to help highlight this: You can learn more about configuring the availability of your virtual machines here. Configuring SQL Server Always On SQL Server Always On is a great feature that you can use with Windows Azure to enable high availability and DR scenarios with SQL Server. Today’s Windows Azure release makes it even easier to configure SQL Server Always On by enabling “Direct Server Return” endpoints to be configured and managed within the Windows Azure Management Portal.  Previously, setting this up required using PowerShell to complete the endpoint configuration.  Starting today you can enable this simply by checking the “Direct Server Return” checkbox: You can learn more about how to use direct server return for SQL Server AlwaysOn availability groups here. Active Directory: Application Access Enhancements This summer we released our initial preview of our Application Access Enhancements for Windows Azure Active Directory.  This service enables you to securely implement single-sign-on (SSO) support against SaaS applications (including Office 365, SalesForce, Workday, Box, Google Apps, GitHub, etc) as well as LOB based applications (including ones built with the new Windows Azure AD support we shipped last week with ASP.NET and VS 2013). Since the initial preview we’ve enhanced our SAML federation capabilities, integrated our new password vaulting system, and shipped multi-factor authentication support. We've also turned on our outbound identity provisioning system and have it working with hundreds of additional SaaS Applications: Earlier this month we published an update on dates and pricing for when the service will be released in general availability form.  In this blog post we announced our intention to release the service in general availability form by the end of the year.  We also announced that the below features would be available in a free tier with it: SSO to every SaaS app we integrate with – Users can Single Sign On to any app we are integrated with at no charge. This includes all the top SAAS Apps and every app in our application gallery whether they use federation or password vaulting. Application access assignment and removal – IT Admins can assign access privileges to web applications to the users in their active directory assuring that every employee has access to the SAAS Apps they need. And when a user leaves the company or changes jobs, the admin can just as easily remove their access privileges assuring data security and minimizing IP loss User provisioning (and de-provisioning) – IT admins will be able to automatically provision users in 3rd party SaaS applications like Box, Salesforce.com, GoToMeeting, DropBox and others. We are working with key partners in the ecosystem to establish these connections, meaning you no longer have to continually update user records in multiple systems. Security and auditing reports – Security is a key priority for us. With the free version of these enhancements you'll get access to our standard set of access reports giving you visibility into which users are using which applications, when they were using them and where they are using them from. In addition, we'll alert you to un-usual usage patterns for instance when a user logs in from multiple locations at the same time. Our Application Access Panel – Users are logging in from every type of devices including Windows, iOS, & Android. Not all of these devices handle authentication in the same manner but the user doesn't care. They need to access their apps from the devices they love. Our Application Access Panel will support the ability for users to access access and launch their apps from any device and anywhere. You can learn more about our plans for application management with Windows Azure Active Directory here.  Try out the preview and start using it today. Enterprise Management: Use Active Directory to Better Manage Windows Azure Windows Azure Active Directory provides the ability to manage your organization in a directory which is hosted entirely in the cloud, or alternatively kept in sync with an on-premises Windows Server Active Directory solution (allowing you to seamlessly integrate with the directory you already have).  With today’s Windows Azure release we are integrating Windows Azure Active Directory even more within the core Windows Azure management experience, and enabling an even richer enterprise security offering.  Specifically: 1) All Windows Azure accounts now have a default Windows Azure Active Directory created for them.  You can create and map any users you want into this directory, and grant administrative rights to manage resources in Windows Azure to these users. 2) You can keep this directory entirely hosted in the cloud – or optionally sync it with your on-premises Windows Server Active Directory.  Both options are free.  The later approach is ideal for companies that wish to use their corporate user identities to sign-in and manage Windows Azure resources.  It also ensures that if an employee leaves an organization, his or her access control rights to the company’s Windows Azure resources are immediately revoked. 3) The Windows Azure Service Management APIs have been updated to support using Windows Azure Active Directory credentials to sign-in and perform management operations.  Prior to today’s release customers had to download and use management certificates (which were not scoped to individual users) to perform management operations.  We still support this management certificate approach (don’t worry – nothing will stop working).  But we think the new Windows Azure Active Directory authentication support enables an even easier and more secure way for customers to manage resources going forward.  4) The Windows Azure SDK 2.2 release (which is also shipping today) includes built-in support for the new Service Management APIs that authenticate with Windows Azure Active Directory, and now allow you to create and manage Windows Azure applications and resources directly within Visual Studio using your Active Directory credentials.  This, combined with updated PowerShell scripts that also support Active Directory, enables an end-to-end enterprise authentication story with Windows Azure. Below are some details on how all of this works: Subscriptions within a Directory As part of today’s update, we have associated all existing Window Azure accounts with a Windows Azure Active Directory (and created one for you if you don’t already have one). When you login to the Windows Azure Management Portal you’ll now see the directory name in the URI of the browser.  For example, in the screen-shot below you can see that I have a “scottgu” directory that my subscriptions are hosted within: Note that you can continue to use Microsoft Accounts (formerly known as Microsoft Live IDs) to sign-into Windows Azure.  These map just fine to a Windows Azure Active Directory – so there is no need to create new usernames that are specific to a directory if you don’t want to.  In the scenario above I’m actually logged in using my @hotmail.com based Microsoft ID which is now mapped to a “scottgu” active directory that was created for me.  By default everything will continue to work just like you used to before. Manage your Directory You can manage an Active Directory (including the one we now create for you by default) by clicking the “Active Directory” tab in the left-hand side of the portal.  This will list all of the directories in your account.  Clicking one the first time will display a getting started page that provides documentation and links to perform common tasks with it: You can use the built-in directory management support within the Windows Azure Management Portal to add/remove/manage users within the directory, enable multi-factor authentication, associate a custom domain (e.g. mycompanyname.com) with the directory, and/or rename the directory to whatever friendly name you want (just click the configure tab to do this).  You can also setup the directory to automatically sync with an on-premises Active Directory using the “Directory Integration” tab. Note that users within a directory by default do not have admin rights to login or manage Windows Azure based resources.  You still need to explicitly grant them co-admin permissions on a subscription for them to login or manage resources in Windows Azure.  You can do this by clicking the Settings tab on the left-hand side of the portal and then by clicking the administrators tab within it. Sign-In Integration within Visual Studio If you install the new Windows Azure SDK 2.2 release, you can now connect to Windows Azure from directly inside Visual Studio without having to download any management certificates.  You can now just right-click on the “Windows Azure” icon within the Server Explorer and choose the “Connect to Windows Azure” context menu option to do so: Doing this will prompt you to enter the email address of the username you wish to sign-in with (make sure this account is a user in your directory with co-admin rights on a subscription): You can use either a Microsoft Account (e.g. Windows Live ID) or an Active Directory based Organizational account as the email.  The dialog will update with an appropriate login prompt depending on which type of email address you enter: Once you sign-in you’ll see the Windows Azure resources that you have permissions to manage show up automatically within the Visual Studio server explorer and be available to start using: No downloading of management certificates required.  All of the authentication was handled using your Windows Azure Active Directory! Manage Subscriptions across Multiple Directories If you have already have multiple directories and multiple subscriptions within your Windows Azure account, we have done our best to create a good default mapping of your subscriptions->directories as part of today’s update.  If you don’t like the default subscription-to-directory mapping we have done you can click the Settings tab in the left-hand navigation of the Windows Azure Management Portal and browse to the Subscriptions tab within it: If you want to map a subscription under a different directory in your account, simply select the subscription from the list, and then click the “Edit Directory” button to choose which directory to map it to.  Mapping a subscription to a different directory takes only seconds and will not cause any of the resources within the subscription to recycle or stop working.  We’ve made the directory->subscription mapping process self-service so that you always have complete control and can map things however you want. Filtering By Directory and Subscription Within the Windows Azure Management Portal you can filter resources in the portal by subscription (allowing you to show/hide different subscriptions).  If you have subscriptions mapped to multiple directory tenants, we also now have a filter drop-down that allows you to filter the subscription list by directory tenant.  This filter is only available if you have multiple subscriptions mapped to multiple directories within your Windows Azure Account:   Windows Azure SDK 2.2 Today we are also releasing a major update of our Windows Azure SDK.  The Windows Azure SDK 2.2 release adds some great new features including: Visual Studio 2013 Support Integrated Windows Azure Sign-In support within Visual Studio Remote Debugging Cloud Services with Visual Studio Firewall Management support within Visual Studio for SQL Databases Visual Studio 2013 RTM VM Images for MSDN Subscribers Windows Azure Management Libraries for .NET Updated Windows Azure PowerShell Cmdlets and ScriptCenter I’ll post a follow-up blog shortly with more details about all of the above. Additional Updates In addition to the above enhancements, today’s release also includes a number of additional improvements: AutoScale: Richer time and date based scheduling support (set different rules on different dates) AutoScale: Ability to Scale to Zero Virtual Machines (very useful for Dev/Test scenarios) AutoScale: Support for time-based scheduling of Mobile Service AutoScale rules Operation Logs: Auditing support for Service Bus management operations Today we also shipped a major update to the Windows Azure SDK – Windows Azure SDK 2.2.  It has so much goodness in it that I have a whole second blog post coming shortly on it! :-) Summary Today’s Windows Azure release enables a bunch of great new scenarios, and enables a much richer enterprise authentication offering. If you don’t already have a Windows Azure account, you can sign-up for a free trial and start using all of the above features today.  Then visit the Windows Azure Developer Center to learn more about how to build apps with it. 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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  • Scrum in 5 Minutes

    - by Stephen.Walther
    The goal of this blog entry is to explain the basic concepts of Scrum in less than five minutes. You learn how Scrum can help a team of developers to successfully complete a complex software project. Product Backlog and the Product Owner Imagine that you are part of a team which needs to create a new website – for example, an e-commerce website. You have an overwhelming amount of work to do. You need to build (or possibly buy) a shopping cart, install an SSL certificate, create a product catalog, create a Facebook page, and at least a hundred other things that you have not thought of yet. According to Scrum, the first thing you should do is create a list. Place the highest priority items at the top of the list and the lower priority items lower in the list. For example, creating the shopping cart and buying the domain name might be high priority items and creating a Facebook page might be a lower priority item. In Scrum, this list is called the Product Backlog. How do you prioritize the items in the Product Backlog? Different stakeholders in the project might have different priorities. Gary, your division VP, thinks that it is crucial that the e-commerce site has a mobile app. Sally, your direct manager, thinks taking advantage of new HTML5 features is much more important. Multiple people are pulling you in different directions. According to Scrum, it is important that you always designate one person, and only one person, as the Product Owner. The Product Owner is the person who decides what items should be added to the Product Backlog and the priority of the items in the Product Backlog. The Product Owner could be the customer who is paying the bills, the project manager who is responsible for delivering the project, or a customer representative. The critical point is that the Product Owner must always be a single person and that single person has absolute authority over the Product Backlog. Sprints and the Sprint Backlog So now the developer team has a prioritized list of items and they can start work. The team starts implementing the first item in the Backlog — the shopping cart — and the team is making good progress. Unfortunately, however, half-way through the work of implementing the shopping cart, the Product Owner changes his mind. The Product Owner decides that it is much more important to create the product catalog before the shopping cart. With some frustration, the team switches their developmental efforts to focus on implementing the product catalog. However, part way through completing this work, once again the Product Owner changes his mind about the highest priority item. Getting work done when priorities are constantly shifting is frustrating for the developer team and it results in lower productivity. At the same time, however, the Product Owner needs to have absolute authority over the priority of the items which need to get done. Scrum solves this conflict with the concept of Sprints. In Scrum, a developer team works in Sprints. At the beginning of a Sprint the developers and the Product Owner agree on the items from the backlog which they will complete during the Sprint. This subset of items from the Product Backlog becomes the Sprint Backlog. During the Sprint, the Product Owner is not allowed to change the items in the Sprint Backlog. In other words, the Product Owner cannot shift priorities on the developer team during the Sprint. Different teams use Sprints of different lengths such as one month Sprints, two-week Sprints, and one week Sprints. For high-stress, time critical projects, teams typically choose shorter sprints such as one week sprints. For more mature projects, longer one month sprints might be more appropriate. A team can pick whatever Sprint length makes sense for them just as long as the team is consistent. You should pick a Sprint length and stick with it. Daily Scrum During a Sprint, the developer team needs to have meetings to coordinate their work on completing the items in the Sprint Backlog. For example, the team needs to discuss who is working on what and whether any blocking issues have been discovered. Developers hate meetings (well, sane developers hate meetings). Meetings take developers away from their work of actually implementing stuff as opposed to talking about implementing stuff. However, a developer team which never has meetings and never coordinates their work also has problems. For example, Fred might get stuck on a programming problem for days and never reach out for help even though Tom (who sits in the cubicle next to him) has already solved the very same problem. Or, both Ted and Fred might have started working on the same item from the Sprint Backlog at the same time. In Scrum, these conflicting needs – limiting meetings but enabling team coordination – are resolved with the idea of the Daily Scrum. The Daily Scrum is a meeting for coordinating the work of the developer team which happens once a day. To keep the meeting short, each developer answers only the following three questions: 1. What have you done since yesterday? 2. What do you plan to do today? 3. Any impediments in your way? During the Daily Scrum, developers are not allowed to talk about issues with their cat, do demos of their latest work, or tell heroic stories of programming problems overcome. The meeting must be kept short — typically about 15 minutes. Issues which come up during the Daily Scrum should be discussed in separate meetings which do not involve the whole developer team. Stories and Tasks Items in the Product or Sprint Backlog – such as building a shopping cart or creating a Facebook page – are often referred to as User Stories or Stories. The Stories are created by the Product Owner and should represent some business need. Unlike the Product Owner, the developer team needs to think about how a Story should be implemented. At the beginning of a Sprint, the developer team takes the Stories from the Sprint Backlog and breaks the stories into tasks. For example, the developer team might take the Create a Shopping Cart story and break it into the following tasks: · Enable users to add and remote items from shopping cart · Persist the shopping cart to database between visits · Redirect user to checkout page when Checkout button is clicked During the Daily Scrum, members of the developer team volunteer to complete the tasks required to implement the next Story in the Sprint Backlog. When a developer talks about what he did yesterday or plans to do tomorrow then the developer should be referring to a task. Stories are owned by the Product Owner and a story is all about business value. In contrast, the tasks are owned by the developer team and a task is all about implementation details. A story might take several days or weeks to complete. A task is something which a developer can complete in less than a day. Some teams get lazy about breaking stories into tasks. Neglecting to break stories into tasks can lead to “Never Ending Stories” If you don’t break a story into tasks, then you can’t know how much of a story has actually been completed because you don’t have a clear idea about the implementation steps required to complete the story. Scrumboard During the Daily Scrum, the developer team uses a Scrumboard to coordinate their work. A Scrumboard contains a list of the stories for the current Sprint, the tasks associated with each Story, and the state of each task. The developer team uses the Scrumboard so everyone on the team can see, at a glance, what everyone is working on. As a developer works on a task, the task moves from state to state and the state of the task is updated on the Scrumboard. Common task states are ToDo, In Progress, and Done. Some teams include additional task states such as Needs Review or Needs Testing. Some teams use a physical Scrumboard. In that case, you use index cards to represent the stories and the tasks and you tack the index cards onto a physical board. Using a physical Scrumboard has several disadvantages. A physical Scrumboard does not work well with a distributed team – for example, it is hard to share the same physical Scrumboard between Boston and Seattle. Also, generating reports from a physical Scrumboard is more difficult than generating reports from an online Scrumboard. Estimating Stories and Tasks Stakeholders in a project, the people investing in a project, need to have an idea of how a project is progressing and when the project will be completed. For example, if you are investing in creating an e-commerce site, you need to know when the site can be launched. It is not enough to just say that “the project will be done when it is done” because the stakeholders almost certainly have a limited budget to devote to the project. The people investing in the project cannot determine the business value of the project unless they can have an estimate of how long it will take to complete the project. Developers hate to give estimates. The reason that developers hate to give estimates is that the estimates are almost always completely made up. For example, you really don’t know how long it takes to build a shopping cart until you finish building a shopping cart, and at that point, the estimate is no longer useful. The problem is that writing code is much more like Finding a Cure for Cancer than Building a Brick Wall. Building a brick wall is very straightforward. After you learn how to add one brick to a wall, you understand everything that is involved in adding a brick to a wall. There is no additional research required and no surprises. If, on the other hand, I assembled a team of scientists and asked them to find a cure for cancer, and estimate exactly how long it will take, they would have no idea. The problem is that there are too many unknowns. I don’t know how to cure cancer, I need to do a lot of research here, so I cannot even begin to estimate how long it will take. So developers hate to provide estimates, but the Product Owner and other product stakeholders, have a legitimate need for estimates. Scrum resolves this conflict by using the idea of Story Points. Different teams use different units to represent Story Points. For example, some teams use shirt sizes such as Small, Medium, Large, and X-Large. Some teams prefer to use Coffee Cup sizes such as Tall, Short, and Grande. Finally, some teams like to use numbers from the Fibonacci series. These alternative units are converted into a Story Point value. Regardless of the type of unit which you use to represent Story Points, the goal is the same. Instead of attempting to estimate a Story in hours (which is doomed to failure), you use a much less fine-grained measure of work. A developer team is much more likely to be able to estimate that a Story is Small or X-Large than the exact number of hours required to complete the story. So you can think of Story Points as a compromise between the needs of the Product Owner and the developer team. When a Sprint starts, the developer team devotes more time to thinking about the Stories in a Sprint and the developer team breaks the Stories into Tasks. In Scrum, you estimate the work required to complete a Story by using Story Points and you estimate the work required to complete a task by using hours. The difference between Stories and Tasks is that you don’t create a task until you are just about ready to start working on a task. A task is something that you should be able to create within a day, so you have a much better chance of providing an accurate estimate of the work required to complete a task than a story. Burndown Charts In Scrum, you use Burndown charts to represent the remaining work on a project. You use Release Burndown charts to represent the overall remaining work for a project and you use Sprint Burndown charts to represent the overall remaining work for a particular Sprint. You create a Release Burndown chart by calculating the remaining number of uncompleted Story Points for the entire Product Backlog every day. The vertical axis represents Story Points and the horizontal axis represents time. A Sprint Burndown chart is similar to a Release Burndown chart, but it focuses on the remaining work for a particular Sprint. There are two different types of Sprint Burndown charts. You can either represent the remaining work in a Sprint with Story Points or with task hours (the following image, taken from Wikipedia, uses hours). When each Product Backlog Story is completed, the Release Burndown chart slopes down. When each Story or task is completed, the Sprint Burndown chart slopes down. Burndown charts typically do not always slope down over time. As new work is added to the Product Backlog, the Release Burndown chart slopes up. If new tasks are discovered during a Sprint, the Sprint Burndown chart will also slope up. The purpose of a Burndown chart is to give you a way to track team progress over time. If, halfway through a Sprint, the Sprint Burndown chart is still climbing a hill then you know that you are in trouble. Team Velocity Stakeholders in a project always want more work done faster. For example, the Product Owner for the e-commerce site wants the website to launch before tomorrow. Developers tend to be overly optimistic. Rarely do developers acknowledge the physical limitations of reality. So Project stakeholders and the developer team often collude to delude themselves about how much work can be done and how quickly. Too many software projects begin in a state of optimism and end in frustration as deadlines zoom by. In Scrum, this problem is overcome by calculating a number called the Team Velocity. The Team Velocity is a measure of the average number of Story Points which a team has completed in previous Sprints. Knowing the Team Velocity is important during the Sprint Planning meeting when the Product Owner and the developer team work together to determine the number of stories which can be completed in the next Sprint. If you know the Team Velocity then you can avoid committing to do more work than the team has been able to accomplish in the past, and your team is much more likely to complete all of the work required for the next Sprint. Scrum Master There are three roles in Scrum: the Product Owner, the developer team, and the Scrum Master. I’v e already discussed the Product Owner. The Product Owner is the one and only person who maintains the Product Backlog and prioritizes the stories. I’ve also described the role of the developer team. The members of the developer team do the work of implementing the stories by breaking the stories into tasks. The final role, which I have not discussed, is the role of the Scrum Master. The Scrum Master is responsible for ensuring that the team is following the Scrum process. For example, the Scrum Master is responsible for making sure that there is a Daily Scrum meeting and that everyone answers the standard three questions. The Scrum Master is also responsible for removing (non-technical) impediments which the team might encounter. For example, if the team cannot start work until everyone installs the latest version of Microsoft Visual Studio then the Scrum Master has the responsibility of working with management to get the latest version of Visual Studio as quickly as possible. The Scrum Master can be a member of the developer team. Furthermore, different people can take on the role of the Scrum Master over time. The Scrum Master, however, cannot be the same person as the Product Owner. Using SonicAgile SonicAgile (SonicAgile.com) is an online tool which you can use to manage your projects using Scrum. You can use the SonicAgile Product Backlog to create a prioritized list of stories. You can estimate the size of the Stories using different Story Point units such as Shirt Sizes and Coffee Cup sizes. You can use SonicAgile during the Sprint Planning meeting to select the Stories that you want to complete during a particular Sprint. You can configure Sprints to be any length of time. SonicAgile calculates Team Velocity automatically and displays a warning when you add too many stories to a Sprint. In other words, it warns you when it thinks you are overcommitting in a Sprint. SonicAgile also includes a Scrumboard which displays the list of Stories selected for a Sprint and the tasks associated with each story. You can drag tasks from one task state to another. Finally, SonicAgile enables you to generate Release Burndown and Sprint Burndown charts. You can use these charts to view the progress of your team. To learn more about SonicAgile, visit SonicAgile.com. Summary In this post, I described many of the basic concepts of Scrum. You learned how a Product Owner uses a Product Backlog to create a prioritized list of tasks. I explained why work is completed in Sprints so the developer team can be more productive. I also explained how a developer team uses the daily scrum to coordinate their work. You learned how the developer team uses a Scrumboard to see, at a glance, who is working on what and the state of each task. I also discussed Burndown charts. You learned how you can use both Release and Sprint Burndown charts to track team progress in completing a project. Finally, I described the crucial role of the Scrum Master – the person who is responsible for ensuring that the rules of Scrum are being followed. My goal was not to describe all of the concepts of Scrum. This post was intended to be an introductory overview. For a comprehensive explanation of Scrum, I recommend reading Ken Schwaber’s book Agile Project Management with Scrum: http://www.amazon.com/Agile-Project-Management-Microsoft-Professional/dp/073561993X/ref=la_B001H6ODMC_1_1?ie=UTF8&qid=1345224000&sr=1-1

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  • Quick guide to Oracle IRM 11g: Classification design

    - by Simon Thorpe
    Quick guide to Oracle IRM 11g indexThis is the final article in the quick guide to Oracle IRM. If you've followed everything prior you will now have a fully functional and tested Information Rights Management service. It doesn't matter if you've been following the 10g or 11g guide as this next article is common to both. ContentsWhy this is the most important part... Understanding the classification and standard rights model Identifying business use cases Creating an effective IRM classification modelOne single classification across the entire businessA context for each and every possible granular use caseWhat makes a good context? Deciding on the use of roles in the context Reviewing the features and security for context roles Summary Why this is the most important part...Now the real work begins, installing and getting an IRM system running is as simple as following instructions. However to actually have an IRM technology easily protecting your most sensitive information without interfering with your users existing daily work flows and be able to scale IRM across the entire business, requires thought into how confidential documents are created, used and distributed. This article is going to give you the information you need to ask the business the right questions so that you can deploy your IRM service successfully. The IRM team here at Oracle have over 10 years of experience in helping customers and it is important you understand the following to be successful in securing access to your most confidential information. Whatever you are trying to secure, be it mergers and acquisitions information, engineering intellectual property, health care documentation or financial reports. No matter what type of user is going to access the information, be they employees, contractors or customers, there are common goals you are always trying to achieve.Securing the content at the earliest point possible and do it automatically. Removing the dependency on the user to decide to secure the content reduces the risk of mistakes significantly and therefore results a more secure deployment. K.I.S.S. (Keep It Simple Stupid) Reduce complexity in the rights/classification model. Oracle IRM lets you make changes to access to documents even after they are secured which allows you to start with a simple model and then introduce complexity once you've understood how the technology is going to be used in the business. After an initial learning period you can review your implementation and start to make informed decisions based on user feedback and administration experience. Clearly communicate to the user, when appropriate, any changes to their existing work practice. You must make every effort to make the transition to sealed content as simple as possible. For external users you must help them understand why you are securing the documents and inform them the value of the technology to both your business and them. Before getting into the detail, I must pay homage to Martin White, Vice President of client services in SealedMedia, the company Oracle acquired and who created Oracle IRM. In the SealedMedia years Martin was involved with every single customer and was key to the design of certain aspects of the IRM technology, specifically the context model we will be discussing here. Listening carefully to customers and understanding the flexibility of the IRM technology, Martin taught me all the skills of helping customers build scalable, effective and simple to use IRM deployments. No matter how well the engineering department designed the software, badly designed and poorly executed projects can result in difficult to use and manage, and ultimately insecure solutions. The advice and information that follows was born with Martin and he's still delivering IRM consulting with customers and can be found at www.thinkers.co.uk. It is from Martin and others that Oracle not only has the most advanced, scalable and usable document security solution on the market, but Oracle and their partners have the most experience in delivering successful document security solutions. Understanding the classification and standard rights model The goal of any successful IRM deployment is to balance the increase in security the technology brings without over complicating the way people use secured content and avoid a significant increase in administration and maintenance. With Oracle it is possible to automate the protection of content, deploy the desktop software transparently and use authentication methods such that users can open newly secured content initially unaware the document is any different to an insecure one. That is until of course they attempt to do something for which they don't have any rights, such as copy and paste to an insecure application or try and print. Central to achieving this objective is creating a classification model that is simple to understand and use but also provides the right level of complexity to meet the business needs. In Oracle IRM the term used for each classification is a "context". A context defines the relationship between.A group of related documents The people that use the documents The roles that these people perform The rights that these people need to perform their role The context is the key to the success of Oracle IRM. It provides the separation of the role and rights of a user from the content itself. Documents are sealed to contexts but none of the rights, user or group information is stored within the content itself. Sealing only places information about the location of the IRM server that sealed it, the context applied to the document and a few other pieces of metadata that pertain only to the document. This important separation of rights from content means that millions of documents can be secured against a single classification and a user needs only one right assigned to be able to access all documents. If you have followed all the previous articles in this guide, you will be ready to start defining contexts to which your sensitive information will be protected. But before you even start with IRM, you need to understand how your own business uses and creates sensitive documents and emails. Identifying business use cases Oracle is able to support multiple classification systems, but usually there is one single initial need for the technology which drives a deployment. This need might be to protect sensitive mergers and acquisitions information, engineering intellectual property, financial documents. For this and every subsequent use case you must understand how users create and work with documents, to who they are distributed and how the recipients should interact with them. A successful IRM deployment should start with one well identified use case (we go through some examples towards the end of this article) and then after letting this use case play out in the business, you learn how your users work with content, how well your communication to the business worked and if the classification system you deployed delivered the right balance. It is at this point you can start rolling the technology out further. Creating an effective IRM classification model Once you have selected the initial use case you will address with IRM, you need to design a classification model that defines the access to secured documents within the use case. In Oracle IRM there is an inbuilt classification system called the "context" model. In Oracle IRM 11g it is possible to extend the server to support any rights classification model, but the majority of users who are not using an application integration (such as Oracle IRM within Oracle Beehive) are likely to be starting out with the built in context model. Before looking at creating a classification system with IRM, it is worth reviewing some recognized standards and methods for creating and implementing security policy. A very useful set of documents are the ISO 17799 guidelines and the SANS security policy templates. First task is to create a context against which documents are to be secured. A context consists of a group of related documents (all top secret engineering research), a list of roles (contributors and readers) which define how users can access documents and a list of users (research engineers) who have been given a role allowing them to interact with sealed content. Before even creating the first context it is wise to decide on a philosophy which will dictate the level of granularity, the question is, where do you start? At a department level? By project? By technology? First consider the two ends of the spectrum... One single classification across the entire business Imagine that instead of having separate contexts, one for engineering intellectual property, one for your financial data, one for human resources personally identifiable information, you create one context for all documents across the entire business. Whilst you may have immediate objections, there are some significant benefits in thinking about considering this. Document security classification decisions are simple. You only have one context to chose from! User provisioning is simple, just make sure everyone has a role in the only context in the business. Administration is very low, if you assign rights to groups from the business user repository you probably never have to touch IRM administration again. There are however some obvious downsides to this model.All users in have access to all IRM secured content. So potentially a sales person could access sensitive mergers and acquisition documents, if they can get their hands on a copy that is. You cannot delegate control of different documents to different parts of the business, this may not satisfy your regulatory requirements for the separation and delegation of duties. Changing a users role affects every single document ever secured. Even though it is very unlikely a business would ever use one single context to secure all their sensitive information, thinking about this scenario raises one very important point. Just having one single context and securing all confidential documents to it, whilst incurring some of the problems detailed above, has one huge value. Once secured, IRM protected content can ONLY be accessed by authorized users. Just think of all the sensitive documents in your business today, imagine if you could ensure that only everyone you trust could open them. Even if an employee lost a laptop or someone accidentally sent an email to the wrong recipient, only the right people could open that file. A context for each and every possible granular use case Now let's think about the total opposite of a single context design. What if you created a context for each and every single defined business need and created multiple contexts within this for each level of granularity? Let's take a use case where we need to protect engineering intellectual property. Imagine we have 6 different engineering groups, and in each we have a research department, a design department and manufacturing. The company information security policy defines 3 levels of information sensitivity... restricted, confidential and top secret. Then let's say that each group and department needs to define access to information from both internal and external users. Finally add into the mix that they want to review the rights model for each context every financial quarter. This would result in a huge amount of contexts. For example, lets just look at the resulting contexts for one engineering group. Q1FY2010 Restricted Internal - Engineering Group 1 - Research Q1FY2010 Restricted Internal - Engineering Group 1 - Design Q1FY2010 Restricted Internal - Engineering Group 1 - Manufacturing Q1FY2010 Restricted External- Engineering Group 1 - Research Q1FY2010 Restricted External - Engineering Group 1 - Design Q1FY2010 Restricted External - Engineering Group 1 - Manufacturing Q1FY2010 Confidential Internal - Engineering Group 1 - Research Q1FY2010 Confidential Internal - Engineering Group 1 - Design Q1FY2010 Confidential Internal - Engineering Group 1 - Manufacturing Q1FY2010 Confidential External - Engineering Group 1 - Research Q1FY2010 Confidential External - Engineering Group 1 - Design Q1FY2010 Confidential External - Engineering Group 1 - Manufacturing Q1FY2010 Top Secret Internal - Engineering Group 1 - Research Q1FY2010 Top Secret Internal - Engineering Group 1 - Design Q1FY2010 Top Secret Internal - Engineering Group 1 - Manufacturing Q1FY2010 Top Secret External - Engineering Group 1 - Research Q1FY2010 Top Secret External - Engineering Group 1 - Design Q1FY2010 Top Secret External - Engineering Group 1 - Manufacturing Now multiply the above by 6 for each engineering group, 18 contexts. You are then creating/reviewing another 18 every 3 months. After a year you've got 72 contexts. What would be the advantages of such a complex classification model? You can satisfy very granular rights requirements, for example only an authorized engineering group 1 researcher can create a top secret report for access internally, and his role will be reviewed on a very frequent basis. Your business may have very complex rights requirements and mapping this directly to IRM may be an obvious exercise. The disadvantages of such a classification model are significant...Huge administrative overhead. Someone in the business must manage, review and administrate each of these contexts. If the engineering group had a single administrator, they would have 72 classifications to reside over each year. From an end users perspective life will be very confusing. Imagine if a user has rights in just 6 of these contexts. They may be able to print content from one but not another, be able to edit content in 2 contexts but not the other 4. Such confusion at the end user level causes frustration and resistance to the use of the technology. Increased synchronization complexity. Imagine a user who after 3 years in the company ends up with over 300 rights in many different contexts across the business. This would result in long synchronization times as the client software updates all your offline rights. Hard to understand who can do what with what. Imagine being the VP of engineering and as part of an internal security audit you are asked the question, "What rights to researchers have to our top secret information?". In this complex model the answer is not simple, it would depend on many roles in many contexts. Of course this example is extreme, but it highlights that trying to build many barriers in your business can result in a nightmare of administration and confusion amongst users. In the real world what we need is a balance of the two. We need to seek an optimum number of contexts. Too many contexts are unmanageable and too few contexts does not give fine enough granularity. What makes a good context? Good context design derives mainly from how well you understand your business requirements to secure access to confidential information. Some customers I have worked with can tell me exactly the documents they wish to secure and know exactly who should be opening them. However there are some customers who know only of the government regulation that requires them to control access to certain types of information, they don't actually know where the documents are, how they are created or understand exactly who should have access. Therefore you need to know how to ask the business the right questions that lead to information which help you define a context. First ask these questions about a set of documentsWhat is the topic? Who are legitimate contributors on this topic? Who are the authorized readership? If the answer to any one of these is significantly different, then it probably merits a separate context. Remember that sealed documents are inherently secure and as such they cannot leak to your competitors, therefore it is better sealed to a broad context than not sealed at all. Simplicity is key here. Always revert to the first extreme example of a single classification, then work towards essential complexity. If there is any doubt, always prefer fewer contexts. Remember, Oracle IRM allows you to change your mind later on. You can implement a design now and continue to change and refine as you learn how the technology is used. It is easy to go from a simple model to a more complex one, it is much harder to take a complex model that is already embedded in the work practice of users and try to simplify it. It is also wise to take a single use case and address this first with the business. Don't try and tackle many different problems from the outset. Do one, learn from the process, refine it and then take what you have learned into the next use case, refine and continue. Once you have a good grasp of the technology and understand how your business will use it, you can then start rolling out the technology wider across the business. Deciding on the use of roles in the context Once you have decided on that first initial use case and a context to create let's look at the details you need to decide upon. For each context, identify; Administrative rolesBusiness owner, the person who makes decisions about who may or may not see content in this context. This is often the person who wanted to use IRM and drove the business purchase. They are the usually the person with the most at risk when sensitive information is lost. Point of contact, the person who will handle requests for access to content. Sometimes the same as the business owner, sometimes a trusted secretary or administrator. Context administrator, the person who will enact the decisions of the Business Owner. Sometimes the point of contact, sometimes a trusted IT person. Document related rolesContributors, the people who create and edit documents in this context. Reviewers, the people who are involved in reviewing documents but are not trusted to secure information to this classification. This role is not always necessary. (See later discussion on Published-work and Work-in-Progress) Readers, the people who read documents from this context. Some people may have several of the roles above, which is fine. What you are trying to do is understand and define how the business interacts with your sensitive information. These roles obviously map directly to roles available in Oracle IRM. Reviewing the features and security for context roles At this point we have decided on a classification of information, understand what roles people in the business will play when administrating this classification and how they will interact with content. The final piece of the puzzle in getting the information for our first context is to look at the permissions people will have to sealed documents. First think why are you protecting the documents in the first place? It is to prevent the loss of leaking of information to the wrong people. To control the information, making sure that people only access the latest versions of documents. You are not using Oracle IRM to prevent unauthorized people from doing legitimate work. This is an important point, with IRM you can erect many barriers to prevent access to content yet too many restrictions and authorized users will often find ways to circumvent using the technology and end up distributing unprotected originals. Because IRM is a security technology, it is easy to get carried away restricting different groups. However I would highly recommend starting with a simple solution with few restrictions. Ensure that everyone who reasonably needs to read documents can do so from the outset. Remember that with Oracle IRM you can change rights to content whenever you wish and tighten security. Always return to the fact that the greatest value IRM brings is that ONLY authorized users can access secured content, remember that simple "one context for the entire business" model. At the start of the deployment you really need to aim for user acceptance and therefore a simple model is more likely to succeed. As time passes and users understand how IRM works you can start to introduce more restrictions and complexity. Another key aspect to focus on is handling exceptions. If you decide on a context model where engineering can only access engineering information, and sales can only access sales data. Act quickly when a sales manager needs legitimate access to a set of engineering documents. Having a quick and effective process for permitting other people with legitimate needs to obtain appropriate access will be rewarded with acceptance from the user community. These use cases can often be satisfied by integrating IRM with a good Identity & Access Management technology which simplifies the process of assigning users the correct business roles. The big print issue... Printing is often an issue of contention, users love to print but the business wants to ensure sensitive information remains in the controlled digital world. There are many cases of physical document loss causing a business pain, it is often overlooked that IRM can help with this issue by limiting the ability to generate physical copies of digital content. However it can be hard to maintain a balance between security and usability when it comes to printing. Consider the following points when deciding about whether to give print rights. Oracle IRM sealed documents can contain watermarks that expose information about the user, time and location of access and the classification of the document. This information would reside in the printed copy making it easier to trace who printed it. Printed documents are slower to distribute in comparison to their digital counterparts, so time sensitive information in printed format may present a lower risk. Print activity is audited, therefore you can monitor and react to users abusing print rights. Summary In summary it is important to think carefully about the way you create your context model. As you ask the business these questions you may get a variety of different requirements. There may be special projects that require a context just for sensitive information created during the lifetime of the project. There may be a department that requires all information in the group is secured and you might have a few senior executives who wish to use IRM to exchange a small number of highly sensitive documents with a very small number of people. Oracle IRM, with its very flexible context classification system, can support all of these use cases. The trick is to introducing the complexity to deliver them at the right level. In another article i'm working on I will go through some examples of how Oracle IRM might map to existing business use cases. But for now, this article covers all the important questions you need to get your IRM service deployed and successfully protecting your most sensitive information.

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