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  • Need to capture and store receiver's details via IPN by using Paypal Mass Pay API

    - by Devner
    Hi all, This is a question about Paypal Mass Pay IPN. My platform is PHP & mySQL. All over the Paypal support website, I have found IPN for only payments made. I need an IPN on similar lines for Mass Pay but could not find it. Also tried experimenting with already existing Mass Pay NVP code, but that did not work either. What I am trying to do is that for all the recipients to whom the payment has been successfully sent via Mass Pay, I want to record their email, amount and unique_id in my own database table. If possible, I want to capture the payment status as well, whether it has been a success of failure and based upon the same, I need to do some in house processing. The existing code Mass pay code is below: <?php $environment = 'sandbox'; // or 'beta-sandbox' or 'live' /** * Send HTTP POST Request * * @param string The API method name * @param string The POST Message fields in &name=value pair format * @return array Parsed HTTP Response body */ function PPHttpPost($methodName_, $nvpStr_) { global $environment; // Set up your API credentials, PayPal end point, and API version. $API_UserName = urlencode('my_api_username'); $API_Password = urlencode('my_api_password'); $API_Signature = urlencode('my_api_signature'); $API_Endpoint = "https://api-3t.paypal.com/nvp"; if("sandbox" === $environment || "beta-sandbox" === $environment) { $API_Endpoint = "https://api-3t.$environment.paypal.com/nvp"; } $version = urlencode('51.0'); // Set the curl parameters. $ch = curl_init(); curl_setopt($ch, CURLOPT_URL, $API_Endpoint); curl_setopt($ch, CURLOPT_VERBOSE, 1); // Turn off the server and peer verification (TrustManager Concept). curl_setopt($ch, CURLOPT_SSL_VERIFYPEER, FALSE); curl_setopt($ch, CURLOPT_SSL_VERIFYHOST, FALSE); curl_setopt($ch, CURLOPT_RETURNTRANSFER, 1); curl_setopt($ch, CURLOPT_POST, 1); // Set the API operation, version, and API signature in the request. $nvpreq = "METHOD=$methodName_&VERSION=$version&PWD=$API_Password&USER=$API_UserName&SIGNATURE=$API_Signature$nvpStr_"; // Set the request as a POST FIELD for curl. curl_setopt($ch, CURLOPT_POSTFIELDS, $nvpreq); // Get response from the server. $httpResponse = curl_exec($ch); if(!$httpResponse) { exit("$methodName_ failed: ".curl_error($ch).'('.curl_errno($ch).')'); } // Extract the response details. $httpResponseAr = explode("&", $httpResponse); $httpParsedResponseAr = array(); foreach ($httpResponseAr as $i => $value) { $tmpAr = explode("=", $value); if(sizeof($tmpAr) > 1) { $httpParsedResponseAr[$tmpAr[0]] = $tmpAr[1]; } } if((0 == sizeof($httpParsedResponseAr)) || !array_key_exists('ACK', $httpParsedResponseAr)) { exit("Invalid HTTP Response for POST request($nvpreq) to $API_Endpoint."); } return $httpParsedResponseAr; } // Set request-specific fields. $emailSubject =urlencode('example_email_subject'); $receiverType = urlencode('EmailAddress'); $currency = urlencode('USD'); // or other currency ('GBP', 'EUR', 'JPY', 'CAD', 'AUD') // Add request-specific fields to the request string. $nvpStr="&EMAILSUBJECT=$emailSubject&RECEIVERTYPE=$receiverType&CURRENCYCODE=$currency"; $receiversArray = array(); for($i = 0; $i < 3; $i++) { $receiverData = array( 'receiverEmail' => "[email protected]", 'amount' => "example_amount", 'uniqueID' => "example_unique_id", 'note' => "example_note"); $receiversArray[$i] = $receiverData; } foreach($receiversArray as $i => $receiverData) { $receiverEmail = urlencode($receiverData['receiverEmail']); $amount = urlencode($receiverData['amount']); $uniqueID = urlencode($receiverData['uniqueID']); $note = urlencode($receiverData['note']); $nvpStr .= "&L_EMAIL$i=$receiverEmail&L_Amt$i=$amount&L_UNIQUEID$i=$uniqueID&L_NOTE$i=$note"; } // Execute the API operation; see the PPHttpPost function above. $httpParsedResponseAr = PPHttpPost('MassPay', $nvpStr); if("SUCCESS" == strtoupper($httpParsedResponseAr["ACK"]) || "SUCCESSWITHWARNING" == strtoupper($httpParsedResponseAr["ACK"])) { exit('MassPay Completed Successfully: '.print_r($httpParsedResponseAr, true)); } else { exit('MassPay failed: ' . print_r($httpParsedResponseAr, true)); } ?> In the code above, how and where do I add code to capture the fields that I requested above? Any code indicating the solution is highly appreciated. Thank you very much.

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  • ASP.NET MVC RememberMe(It's large, please don't quit reading. Have explained the problem in detail a

    - by nccsbim071
    After searching a lot i did not get any answers and finally i had to get back to you. Below i am explaining my problem in detail. It's too long, so please don't quit reading. I have explained my problem in simple language. I have been developing an asp.net mvc project. I am using standard ASP.NET roles and membership. Everything is working fine but the remember me functionality doesn't work at all. I am listing all the details of work. Hope you guys can help me out solve this problem. I simply need this: I need user to login to web application. During login they can either login with remember me or without it. If user logs in with remember me, i want browser to remember them for long time, let's say atleast one year or considerably long time. The way they do it in www.dotnetspider.com,www.codeproject.com,www.daniweb.com and many other sites. If user logs in without remember me, then browser should allow access to website for some 20 -30 minutes and after that their session should expire. Their session should also expire when user logs in and shuts down the browser without logging out. Note: I have succesfully implemented above functionality without using standard asp.net roles and membership by creating my own talbes for user and authenticating against my database table, setting cookie and sessions in my other projects. But for this project we starting from the beginning used standard asp.net roles and membership. We thought it will work and after everything was build at the time of testing it just didn't work. and now we cannot replace the existing functionality with standard asp.net roles and membership with my own custom user tables and all the stuff, you understand what i am taling about. Either there is some kind of bug with standard asp.net roles and membership functionality or i have the whole concept of standard asp.net roles and membership wrong. i have stated what i want above. I think it's very simple and reasonable. What i did Login form with username,password and remember me field. My setting in web.config: <authentication mode="Forms"> <forms loginUrl="~/Account/LogOn" timeout="2880"/> </authentication> in My controller action, i have this: FormsAuth.SignIn(userName, rememberMe); public void SignIn(string userName, bool createPersistentCookie) { FormsAuthentication.SetAuthCookie(userName, createPersistentCookie); } Now the problems are following: I have already stated in above section "I simply need this". user can successfully log in to the system. Their session exists for as much minutes as specified in timeout value in web.config. I have also given a sample of my web.config. In my samplem if i set the timeout to 5 minutes,then user session expires after 5 minutes, that's ok. But if user closes the browser and reopen the browser, user can still enter the website without loggin in untill time specified in "timeout" has not passed out. The sliding expiration for timeout value is also working fine. Now if user logs in to the system with remember me checked, user session still expires after 5 minutes. This is not good behaviour, is it?. I mean to say that if user logs in to the system with remember me checked he should be remembered for a long time untill he doesn't logs out of the system or user doesn't manually deletes all the cookies from the browser. If user logs in to the system without remember me checked his session should expire after the timeout period values specified in web.config and also if users closes the browser. The problem is that if user closes the browser and reopens it he can still enter the website without logging in. I search internet a lot on this topic, but i could not get the solution. In the blog post(http://weblogs.asp.net/scottgu/archive/2005/11/08/430011.aspx) made by Scott Gu on exactly the same topic. The users are complaining about the same thing in their comments ut there is no easy solution given in by Mr. Scott. I read it at following places: http://weblogs.asp.net/scottgu/archive/2005/11/08/430011.aspx http://geekswithblogs.net/vivek/archive/2006/09/14/91191.aspx I guess this is a problem of lot's of users. As seem from blog post made by Mr. Scott Gu. Your help will be really appreciated. Thanks in advance.

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  • Does anyone really understand how HFSC scheduling in Linux/BSD works?

    - by Mecki
    I read the original SIGCOMM '97 PostScript paper about HFSC, it is very technically, but I understand the basic concept. Instead of giving a linear service curve (as with pretty much every other scheduling algorithm), you can specify a convex or concave service curve and thus it is possible to decouple bandwidth and delay. However, even though this paper mentions to kind of scheduling algorithms being used (real-time and link-share), it always only mentions ONE curve per scheduling class (the decoupling is done by specifying this curve, only one curve is needed for that). Now HFSC has been implemented for BSD (OpenBSD, FreeBSD, etc.) using the ALTQ scheduling framework and it has been implemented Linux using the TC scheduling framework (part of iproute2). Both implementations added two additional service curves, that were NOT in the original paper! A real-time service curve and an upper-limit service curve. Again, please note that the original paper mentions two scheduling algorithms (real-time and link-share), but in that paper both work with one single service curve. There never have been two independent service curves for either one as you currently find in BSD and Linux. Even worse, some version of ALTQ seems to add an additional queue priority to HSFC (there is no such thing as priority in the original paper either). I found several BSD HowTo's mentioning this priority setting (even though the man page of the latest ALTQ release knows no such parameter for HSFC, so officially it does not even exist). This all makes the HFSC scheduling even more complex than the algorithm described in the original paper and there are tons of tutorials on the Internet that often contradict each other, one claiming the opposite of the other one. This is probably the main reason why nobody really seems to understand how HFSC scheduling really works. Before I can ask my questions, we need a sample setup of some kind. I'll use a very simple one as seen in the image below: Here are some questions I cannot answer because the tutorials contradict each other: What for do I need a real-time curve at all? Assuming A1, A2, B1, B2 are all 128 kbit/s link-share (no real-time curve for either one), then each of those will get 128 kbit/s if the root has 512 kbit/s to distribute (and A and B are both 256 kbit/s of course), right? Why would I additionally give A1 and B1 a real-time curve with 128 kbit/s? What would this be good for? To give those two a higher priority? According to original paper I can give them a higher priority by using a curve, that's what HFSC is all about after all. By giving both classes a curve of [256kbit/s 20ms 128kbit/s] both have twice the priority than A2 and B2 automatically (still only getting 128 kbit/s on average) Does the real-time bandwidth count towards the link-share bandwidth? E.g. if A1 and B1 both only have 64kbit/s real-time and 64kbit/s link-share bandwidth, does that mean once they are served 64kbit/s via real-time, their link-share requirement is satisfied as well (they might get excess bandwidth, but lets ignore that for a second) or does that mean they get another 64 kbit/s via link-share? So does each class has a bandwidth "requirement" of real-time plus link-share? Or does a class only have a higher requirement than the real-time curve if the link-share curve is higher than the real-time curve (current link-share requirement equals specified link-share requirement minus real-time bandwidth already provided to this class)? Is upper limit curve applied to real-time as well, only to link-share, or maybe to both? Some tutorials say one way, some say the other way. Some even claim upper-limit is the maximum for real-time bandwidth + link-share bandwidth? What is the truth? Assuming A2 and B2 are both 128 kbit/s, does it make any difference if A1 and B1 are 128 kbit/s link-share only, or 64 kbit/s real-time and 128 kbit/s link-share, and if so, what difference? If I use the seperate real-time curve to increase priorities of classes, why would I need "curves" at all? Why is not real-time a flat value and link-share also a flat value? Why are both curves? The need for curves is clear in the original paper, because there is only one attribute of that kind per class. But now, having three attributes (real-time, link-share, and upper-limit) what for do I still need curves on each one? Why would I want the curves shape (not average bandwidth, but their slopes) to be different for real-time and link-share traffic? According to the little documentation available, real-time curve values are totally ignored for inner classes (class A and B), they are only applied to leaf classes (A1, A2, B1, B2). If that is true, why does the ALTQ HFSC sample configuration (search for 3.3 Sample configuration) set real-time curves on inner classes and claims that those set the guaranteed rate of those inner classes? Isn't that completely pointless? (note: pshare sets the link-share curve in ALTQ and grate the real-time curve; you can see this in the paragraph above the sample configuration). Some tutorials say the sum of all real-time curves may not be higher than 80% of the line speed, others say it must not be higher than 70% of the line speed. Which one is right or are they maybe both wrong? One tutorial said you shall forget all the theory. No matter how things really work (schedulers and bandwidth distribution), imagine the three curves according to the following "simplified mind model": real-time is the guaranteed bandwidth that this class will always get. link-share is the bandwidth that this class wants to become fully satisfied, but satisfaction cannot be guaranteed. In case there is excess bandwidth, the class might even get offered more bandwidth than necessary to become satisfied, but it may never use more than upper-limit says. For all this to work, the sum of all real-time bandwidths may not be above xx% of the line speed (see question above, the percentage varies). Question: Is this more or less accurate or a total misunderstanding of HSFC? And if assumption above is really accurate, where is prioritization in that model? E.g. every class might have a real-time bandwidth (guaranteed), a link-share bandwidth (not guaranteed) and an maybe an upper-limit, but still some classes have higher priority needs than other classes. In that case I must still prioritize somehow, even among real-time traffic of those classes. Would I prioritize by the slope of the curves? And if so, which curve? The real-time curve? The link-share curve? The upper-limit curve? All of them? Would I give all of them the same slope or each a different one and how to find out the right slope? I still haven't lost hope that there exists at least a hand full of people in this world that really understood HFSC and are able to answer all these questions accurately. And doing so without contradicting each other in the answers would be really nice ;-)

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  • Am I crazy? (How) should I create a jQuery content editor?

    - by Brendon Muir
    Ok, so I created a CMS mainly aimed at Primary Schools. It's getting fairly popular in New Zealand but the one thing I hate with a passion is the largely bad quality of in browser WYSIWYG editors. I've been using KTML (made by InterAKT which was purchased by Adobe a few years ago). In my opinion this editor does a lot of great things (image editing/management, thumbnailing and pretty good content editing). Unfortunately time has had its nasty way with this product and new browsers are beginning to break features and generally degrade the performance of this tool. It's also quite scary basing my livelihood on a defunct product! I've been hunting, in fact I regularly hunt around to see if anything has changed in the WYSIWYG arena. The closest thing I've seen that excites me is the WYSIHAT framework, but they've decided to ignore a pretty relevant editing paradigm which I'm going to outline below. This is the idea for my proposed editor, and I don't know of any existing products that can do this properly: Right, so the traditional model for editing let's say a Page in a CMS is to log into a 'back end' and click edit on the page. This will then load another screen with the editor in it and perhaps a few other fields. More advanced CMS's will maybe have several editing boxes that are for different portions of the page. Anyway, the big problem with this way of doing things is that the user is editing a document outside of the final context it will appear in. In the simplest terms, this means the page template. Many things can be wrong, e.g. the with of the editing area might be different to the width of the actual template area. The height is nearly always fixed because existing editors always seem to use IFRAMES for backward compatibility. And there are plenty of other beefs which I'm sure you're quite aware of if you're in this development area. Here's my editor utopia: You click 'Edit Page': The actual page (with its actual template) displays. Portions of the page have been marked as editable via a class name. You click on one of these areas (in my case it'd just be the big 'body' area in the middle of the template) and a editing bar drops down from the top of the screen with all your standard controls (bold, italic, insert image etc...). Iframes are never used, instead we rely on setting contentEditable to true on the DIV's in question. Firefox 2 and IE6 can go away, let's move on. You can edit the page knowing exactly how it will look when you save it. Because all the styles for this template are loaded, your headings will look correct, everything will be just dandy. Is this such a radical concept? Why are we still content with TinyMCE and that other editor that is too embarrassing to use because it sounds like a swear word!? Let's face the facts: I'm a JavaScript novice. I did once play around in this area using the Javascript Anthology from SitePoint as a guide. It was quite a cool learning experience, but they of course used the IFRAME to make their lives easier. I tried to go a different route and just use contentEditable and even tried to sidestep the native content editing routines (execCommand) and instead wrote my own. They kind of worked but there were always issues. Now we have jQuery, and a few libraries that abstract things like IE's lack of Range support. I'm wondering, am I crazy, or is it actually a good idea to try and build an editor around this editing paradigm using jQuery and relevant plugins to make the job easier? My actual questions: Where would you start? What plugins do you know of that would help the most? Is it worth it, or is there a magical project that already exists that I should join in on? What are the biggest hurdles to overcome in a project like this? Am I crazy? I hope this question has been posted on the right board. I figured it is a technical question as I'm wanting to know specific hurdles and pitfalls to watch out for and also if it is technically feasible with todays technology. Looking forward to hearing peoples thoughts and opinions.

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  • ASP.NET RememberMe(It's large, please don't quit reading. Have explained the problem in detail and s

    - by nccsbim071
    After searching a lot i did not get any answers and finally i had to get back to you. Below i am explaining my problem in detail. It's too long, so please don't quit reading. I have explained my problem in simple language. I have been developing an asp.net mvc project. I am using standard ASP.NET roles and membership. Everything is working fine but the remember me functionality doesn't work at all. I am listing all the details of work. Hope you guys can help me out solve this problem. I simply need this: I need user to login to web application. During login they can either login with remember me or without it. If user logs in with remember me, i want browser to remember them for long time, let's say atleast one year or considerably long time. The way they do it in www.dotnetspider.com,www.codeproject.com,www.daniweb.com and many other sites. If user logs in without remember me, then browser should allow access to website for some 20 -30 minutes and after that their session should expire. Their session should also expire when user logs in and shuts down the browser without logging out. Note: I have succesfully implemented above functionality without using standard asp.net roles and membership by creating my own talbes for user and authenticating against my database table, setting cookie and sessions in my other projects. But for this project we starting from the beginning used standard asp.net roles and membership. We thought it will work and after everything was build at the time of testing it just didn't work. and now we cannot replace the existing functionality with standard asp.net roles and membership with my own custom user tables and all the stuff, you understand what i am taling about. Either there is some kind of bug with standard asp.net roles and membership functionality or i have the whole concept of standard asp.net roles and membership wrong. i have stated what i want above. I think it's very simple and reasonable. What i did Login form with username,password and remember me field. My setting in web.config: in My controller action, i have this: FormsAuth.SignIn(userName, rememberMe); public void SignIn(string userName, bool createPersistentCookie) { FormsAuthentication.SetAuthCookie(userName, createPersistentCookie); } Now the problems are following: I have already stated in above section "I simply need this". user can successfully log in to the system. Their session exists for as much minutes as specified in timeout value in web.config. I have also given a sample of my web.config. In my samplem if i set the timeout to 5 minutes,then user session expires after 5 minutes, that's ok. But if user closes the browser and reopen the browser, user can still enter the website without loggin in untill time specified in "timeout" has not passed out. The sliding expiration for timeout value is also working fine. Now if user logs in to the system with remember me checked, user session still expires after 5 minutes. This is not good behaviour, is it?. I mean to say that if user logs in to the system with remember me checked he should be remembered for a long time untill he doesn't logs out of the system or user doesn't manually deletes all the cookies from the browser. If user logs in to the system without remember me checked his session should expire after the timeout period values specified in web.config and also if users closes the browser. The problem is that if user closes the browser and reopens it he can still enter the website without logging in. I search internet a lot on this topic, but i could not get the solution. In the blog post(http://weblogs.asp.net/scottgu/archive/2005/11/08/430011.aspx) made by Scott Gu on exactly the same topic. The users are complaining about the same thing in their comments ut there is no easy solution given in by Mr. Scott. I read it at following places: http://weblogs.asp.net/scottgu/archive/2005/11/08/430011.aspx http://geekswithblogs.net/vivek/archive/2006/09/14/91191.aspx I guess this is a problem of lot's of users. As seem from blog post made by Mr. Scott Gu. Your help will be really appreciated. Thanks in advance.

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  • How to get javascript object references or reference count?

    - by Tauren
    How to get reference count for an object Is it possible to determine if a javascript object has multiple references to it? Or if it has references besides the one I'm accessing it with? Or even just to get the reference count itself? Can I find this information from javascript itself, or will I need to keep track of my own reference counters. Obviously, there must be at least one reference to it for my code access the object. But what I want to know is if there are any other references to it, or if my code is the only place it is accessed. I'd like to be able to delete the object if nothing else is referencing it. If you know the answer, there is no need to read the rest of this question. Below is just an example to make things more clear. Use Case In my application, I have a Repository object instance called contacts that contains an array of ALL my contacts. There are also multiple Collection object instances, such as friends collection and a coworkers collection. Each collection contains an array with a different set of items from the contacts Repository. Sample Code To make this concept more concrete, consider the code below. Each instance of the Repository object contains a list of all items of a particular type. You might have a repository of Contacts and a separate repository of Events. To keep it simple, you can just get, add, and remove items, and add many via the constructor. var Repository = function(items) { this.items = items || []; } Repository.prototype.get = function(id) { for (var i=0,len=this.items.length; i<len; i++) { if (items[i].id === id) { return this.items[i]; } } } Repository.prototype.add = function(item) { if (toString.call(item) === "[object Array]") { this.items.concat(item); } else { this.items.push(item); } } Repository.prototype.remove = function(id) { for (var i=0,len=this.items.length; i<len; i++) { if (items[i].id === id) { this.removeIndex(i); } } } Repository.prototype.removeIndex = function(index) { if (items[index]) { if (/* items[i] has more than 1 reference to it */) { // Only remove item from repository if nothing else references it this.items.splice(index,1); return; } } } Note the line in remove with the comment. I only want to remove the item from my master repository of objects if no other objects have a reference to the item. Here's Collection: var Collection = function(repo,items) { this.repo = repo; this.items = items || []; } Collection.prototype.remove = function(id) { for (var i=0,len=this.items.length; i<len; i++) { if (items[i].id === id) { // Remove object from this collection this.items.splice(i,1); // Tell repo to remove it (only if no other references to it) repo.removeIndxe(i); return; } } } And then this code uses Repository and Collection: var contactRepo = new Repository([ {id: 1, name: "Joe"}, {id: 2, name: "Jane"}, {id: 3, name: "Tom"}, {id: 4, name: "Jack"}, {id: 5, name: "Sue"} ]); var friends = new Collection( contactRepo, [ contactRepo.get(2), contactRepo.get(4) ] ); var coworkers = new Collection( contactRepo, [ contactRepo.get(1), contactRepo.get(2), contactRepo.get(5) ] ); contactRepo.items; // contains item ids 1, 2, 3, 4, 5 friends.items; // contains item ids 2, 4 coworkers.items; // contains item ids 1, 2, 5 coworkers.remove(2); contactRepo.items; // contains item ids 1, 2, 3, 4, 5 friends.items; // contains item ids 2, 4 coworkers.items; // contains item ids 1, 5 friends.remove(4); contactRepo.items; // contains item ids 1, 2, 3, 5 friends.items; // contains item ids 2 coworkers.items; // contains item ids 1, 5 Notice how coworkers.remove(2) didn't remove id 2 from contactRepo? This is because it was still referenced from friends.items. However, friends.remove(4) causes id 4 to be removed from contactRepo, because no other collection is referring to it. Summary The above is what I want to do. I'm sure there are ways I can do this by keeping track of my own reference counters and such. But if there is a way to do it using javascript's built-in reference management, I'd like to hear about how to use it.

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  • Trying to get these list items to display inline

    - by Joel
    I have several unordered lists that I want to display like this: <ul> <li><img></li> <li><p></li> //inline </ul> //linebreak <ul> <li><img></li> <li><p></li> //inline </ul> ...etc I'm not able to get the li items to be inline with eachother. They are stacking vertically. I have stripped away most styling but still can't figure out what I'm doing wrong: html: <ul class="instrument"> <li class="imagebox"><img src="/images/matepe.jpg" width="247" height="228" alt="Matepe" /></li> <li class="textbox"><p>The matepe is a 24 key instrument that is played by the Kore-Kore people in North-Eastern Zimbabwe and Mozambique. It utilizes four fingers-each playing an individual melody. These melodies also interwieve to create resultant melodies that can be manipulated thru accenting different fingers. The matepe is used in Rattletree as the bridge from the physical world to the spirit world. The matepe is used in the Kore-Kore culture to summon the Mhondoro spirits which are thought to be able to communicate directly with Mwari (God) without the need of an intermediary.</p></li> </ul> <ul class="instrument"> <li class="imagebox"><img src="/images/soprano_little.jpg" border="0" width="247" height="170" alt="Soprano" /></li> <li class="textbox"><p>The highest voice of the Rattletree Marimba orchestra is the Soprano marimba. The soprano is used to whip up the energy on the dancefloor and help people reach ecstatic state with it's high and clear singing voice. The range of these sopranos goes much lower than 'typical' Zimbabwean style sopranos. The sopranos play the range of the right hand of the matepe and go two notes higher and five notes lower. Rattletree uses two sopranos.</p></li> </ul> <ul class="instrument"> <li class="imagebox"><img src="/images/bari_little.jpg" border="0" width="247" height="170" alt="Baritone" /></li> <li class="textbox"><p>The Baritone is the next lower voice in the orchestra. The bari is where the funk is. Generally bubbling over the Bass line, the baritone creates the syncopations and polyrhythms that messes with the 'minds' of the dancers and helps seperate the listener from the physical realm of thought. The range of the baritone covers the full range of the left hand side of the matepe.</p></li> </ul> <ul class="instrument"> <li class="imagebox"><img src="/images/darren_littlebass.jpg" border="0" width="247" height="195" alt="Bass"/><strong>Bass Marimba</strong></li> <li class="textbox"><p>The towering Bass Marimba is the foundation of the Rattletree Marimba sound. Putting out frequencies as low as 22hZ, the bass creates the drive that gets the dancefloor moving. It is 5.5' tall, 9' long, and 4' deep. It is played by standing on a platform and struck with mallets that have lacross-ball size heads (they are actually made with rubber dog balls). The Bass marimba's range covers the lowest five notes of the matepe and goes another five notes lower.</p></li> </ul> <ul class="instrument"> <li class="imagebox"><img src="/images/wayne_little.jpg" border="0" width="247" height="177" alt="Drums"/><strong>Drumset</strong></li> <li class="textbox"><p>All the intricate polyrhythms are held together tastefully with the drumset. The drums provides the consistancy and grounding that the dancers need to keep going all night. While the steady kick and high-hat provide that grounding function, the toms and snare and allowed to be another voice in the poylrhythmic texture-helping the dancers abandon the concept of a "one" within this cyclical music.</p></li> </ul> css: ul.instrument { text-align:left; display:inline; } ul.instrument li { list-style-type: none; } li.imagebox { } li.textbox { } li.textbox p{ width: 247px; }

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  • Modelling boost::Lockable with semaphore rather than mutex (previously titled: Unlocking a mutex fr

    - by dan
    I'm using the C++ boost::thread library, which in my case means I'm using pthreads. Officially, a mutex must be unlocked from the same thread which locks it, and I want the effect of being able to lock in one thread and then unlock in another. There are many ways to accomplish this. One possibility would be to write a new mutex class which allows this behavior. For example: class inter_thread_mutex{ bool locked; boost::mutex mx; boost::condition_variable cv; public: void lock(){ boost::unique_lock<boost::mutex> lck(mx); while(locked) cv.wait(lck); locked=true; } void unlock(){ { boost::lock_guard<boost::mutex> lck(mx); if(!locked) error(); locked=false; } cv.notify_one(); } // bool try_lock(); void error(); etc. } I should point out that the above code doesn't guarantee FIFO access, since if one thread calls lock() while another calls unlock(), this first thread may acquire the lock ahead of other threads which are waiting. (Come to think of it, the boost::thread documentation doesn't appear to make any explicit scheduling guarantees for either mutexes or condition variables). But let's just ignore that (and any other bugs) for now. My question is, if I decide to go this route, would I be able to use such a mutex as a model for the boost Lockable concept. For example, would anything go wrong if I use a boost::unique_lock< inter_thread_mutex for RAII-style access, and then pass this lock to boost::condition_variable_any.wait(), etc. On one hand I don't see why not. On the other hand, "I don't see why not" is usually a very bad way of determining whether something will work. The reason I ask is that if it turns out that I have to write wrapper classes for RAII locks and condition variables and whatever else, then I'd rather just find some other way to achieve the same effect. EDIT: The kind of behavior I want is basically as follows. I have an object, and it needs to be locked whenever it is modified. I want to lock the object from one thread, and do some work on it. Then I want to keep the object locked while I tell another worker thread to complete the work. So the first thread can go on and do something else while the worker thread finishes up. When the worker thread gets done, it unlocks the mutex. And I want the transition to be seemless so nobody else can get the mutex lock in between when thread 1 starts the work and thread 2 completes it. Something like inter_thread_mutex seems like it would work, and it would also allow the program to interact with it as if it were an ordinary mutex. So it seems like a clean solution. If there's a better solution, I'd be happy to hear that also. EDIT AGAIN: The reason I need locks to begin with is that there are multiple master threads, and the locks are there to prevent them from accessing shared objects concurrently in invalid ways. So the code already uses loop-level lock-free sequencing of operations at the master thread level. Also, in the original implementation, there were no worker threads, and the mutexes were ordinary kosher mutexes. The inter_thread_thingy came up as an optimization, primarily to improve response time. In many cases, it was sufficient to guarantee that the "first part" of operation A, occurs before the "first part" of operation B. As a dumb example, say I punch object 1 and give it a black eye. Then I tell object 1 to change it's internal structure to reflect all the tissue damage. I don't want to wait around for the tissue damage before I move on to punch object 2. However, I do want the tissue damage to occur as part of the same operation; for example, in the interim, I don't want any other thread to reconfigure the object in such a way that would make tissue damage an invalid operation. (yes, this example is imperfect in many ways, and no I'm not working on a game) So we made the change to a model where ownership of an object can be passed to a worker thread to complete an operation, and it actually works quite nicely; each master thread is able to get a lot more operations done because it doesn't need to wait for them all to complete. And, since the event sequencing at the master thread level is still loop-based, it is easy to write high-level master-thread operations, as they can be based on the assumption that an operation is complete when the corresponding function call returns. Finally, I thought it would be nice to use inter_thread mutex/semaphore thingies using RAII with boost locks to encapsulate the necessary synchronization that is required to make the whole thing work.

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  • Graphics.MeasureCharacterRanges giving wrong size calculations in C#.Net?

    - by Owen Blacker
    I'm trying to render some text into a specific part of an image in a Web Forms app. The text will be user entered, so I want to vary the font size to make sure it fits within the bounding box. I have code that was doing this fine on my proof-of-concept implementation, but I'm now trying it against the assets from the designer, which are larger, and I'm getting some odd results. I'm running the size calculation as follows: StringFormat fmt = new StringFormat(); fmt.Alignment = StringAlignment.Center; fmt.LineAlignment = StringAlignment.Near; fmt.FormatFlags = StringFormatFlags.NoClip; fmt.Trimming = StringTrimming.None; int size = __startingSize; Font font = __fonts.GetFontBySize(size); while (GetStringBounds(text, font, fmt).IsLargerThan(__textBoundingBox)) { context.Trace.Write("MyHandler.ProcessRequest", "Decrementing font size to " + size + ", as size is " + GetStringBounds(text, font, fmt).Size() + " and limit is " + __textBoundingBox.Size()); size--; if (size < __minimumSize) { break; } font = __fonts.GetFontBySize(size); } context.Trace.Write("MyHandler.ProcessRequest", "Writing " + text + " in " + font.FontFamily.Name + " at " + font.SizeInPoints + "pt, size is " + GetStringBounds(text, font, fmt).Size() + " and limit is " + __textBoundingBox.Size()); I then use the following line to render the text onto an image I'm pulling from the filesystem: g.DrawString(text, font, __brush, __textBoundingBox, fmt); where: __fonts is a PrivateFontCollection, PrivateFontCollection.GetFontBySize is an extension method that returns a FontFamily RectangleF __textBoundingBox = new RectangleF(150, 110, 212, 64); int __minimumSize = 8; int __startingSize = 48; Brush __brush = Brushes.White; int size starts out at 48 and decrements within that loop Graphics g has SmoothingMode.AntiAlias and TextRenderingHint.AntiAlias set context is a System.Web.HttpContext (this is an excerpt from the ProcessRequest method of an IHttpHandler) The other methods are: private static RectangleF GetStringBounds(string text, Font font, StringFormat fmt) { CharacterRange[] range = { new CharacterRange(0, text.Length) }; StringFormat myFormat = fmt.Clone() as StringFormat; myFormat.SetMeasurableCharacterRanges(range); using (Graphics g = Graphics.FromImage(new Bitmap( (int) __textBoundingBox.Width - 1, (int) __textBoundingBox.Height - 1))) { g.SmoothingMode = System.Drawing.Drawing2D.SmoothingMode.AntiAlias; g.TextRenderingHint = System.Drawing.Text.TextRenderingHint.AntiAlias; Region[] regions = g.MeasureCharacterRanges(text, font, __textBoundingBox, myFormat); return regions[0].GetBounds(g); } } public static string Size(this RectangleF rect) { return rect.Width + "×" + rect.Height; } public static bool IsLargerThan(this RectangleF a, RectangleF b) { return (a.Width > b.Width) || (a.Height > b.Height); } Now I have two problems. Firstly, the text sometimes insists on wrapping by inserting a line-break within a word, when it should just fail to fit and cause the while loop to decrement again. I can't see why it is that Graphics.MeasureCharacterRanges thinks that this fits within the box when it shouldn't be word-wrapping within a word. This behaviour is exhibited irrespective of the character set used (I get it in Latin alphabet words, as well as other parts of the Unicode range, like Cyrillic, Greek, Georgian and Armenian). Is there some setting I should be using to force Graphics.MeasureCharacterRanges only to be word-wrapping at whitespace characters (or hyphens)? This first problem is the same as post 2499067. Secondly, in scaling up to the new image and font size, Graphics.MeasureCharacterRanges is giving me heights that are wildly off. The RectangleF I am drawing within corresponds to a visually apparent area of the image, so I can easily see when the text is being decremented more than is necessary. Yet when I pass it some text, the GetBounds call is giving me a height that is almost double what it's actually taking. Using trial and error to set the __minimumSize to force an exit from the while loop, I can see that 24pt text fits within the bounding box, yet Graphics.MeasureCharacterRanges is reporting that the height of that text, once rendered to the image, is 122px (when the bounding box is 64px tall and it fits within that box). Indeed, without forcing the matter, the while loop iterates to 18pt, at which point Graphics.MeasureCharacterRanges returns a value that fits. The trace log excerpt is as follows: Decrementing font size to 24, as size is 193×122 and limit is 212×64 Decrementing font size to 23, as size is 191×117 and limit is 212×64 Decrementing font size to 22, as size is 200×75 and limit is 212×64 Decrementing font size to 21, as size is 192×71 and limit is 212×64 Decrementing font size to 20, as size is 198×68 and limit is 212×64 Decrementing font size to 19, as size is 185×65 and limit is 212×64 Writing VENNEGOOR of HESSELINK in DIN-Black at 18pt, size is 178×61 and limit is 212×64 So why is Graphics.MeasureCharacterRanges giving me a wrong result? I could understand it being, say, the line height of the font if the loop stopped around 21pt (which would visually fit, if I screenshot the results and measure it in Paint.Net), but it's going far further than it should be doing because, frankly, it's returning the wrong damn results. Any and all help gratefully received. Thanks!

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  • What does Ruby have that Python doesn't, and vice versa?

    - by Lennart Regebro
    There is a lot of discussions of Python vs Ruby, and I all find them completely unhelpful, because they all turn around why feature X sucks in language Y, or that claim language Y doesn't have X, although in fact it does. I also know exactly why I prefer Python, but that's also subjective, and wouldn't help anybody choosing, as they might not have the same tastes in development as I do. It would therefore be interesting to list the differences, objectively. So no "Python's lambdas sucks". Instead explain what Ruby's lambdas can do that Python's can't. No subjectivity. Example code is good! Don't have several differences in one answer, please. And vote up the ones you know are correct, and down those you know are incorrect (or are subjective). Also, differences in syntax is not interesting. We know Python does with indentation what Ruby does with brackets and ends, and that @ is called self in Python. UPDATE: This is now a community wiki, so we can add the big differences here. Ruby has a class reference in the class body In Ruby you have a reference to the class (self) already in the class body. In Python you don't have a reference to the class until after the class construction is finished. An example: class Kaka puts self end self in this case is the class, and this code would print out "Kaka". There is no way to print out the class name or in other ways access the class from the class definition body in Python. All classes are mutable in Ruby This lets you develop extensions to core classes. Here's an example of a rails extension: class String def starts_with?(other) head = self[0, other.length] head == other end end Ruby has Perl-like scripting features Ruby has first class regexps, $-variables, the awk/perl line by line input loop and other features that make it more suited to writing small shell scripts that munge text files or act as glue code for other programs. Ruby has first class continuations Thanks to the callcc statement. In Python you can create continuations by various techniques, but there is no support built in to the language. Ruby has blocks With the "do" statement you can create a multi-line anonymous function in Ruby, which will be passed in as an argument into the method in front of do, and called from there. In Python you would instead do this either by passing a method or with generators. Ruby: amethod { |here| many=lines+of+code goes(here) } Python: def function(here): many=lines+of+code goes(here) amethod(function) Interestingly, the convenience statement in Ruby for calling a block is called "yield", which in Python will create a generator. Ruby: def themethod yield 5 end themethod do |foo| puts foo end Python: def themethod(): yield 5 for foo in themethod: print foo Although the principles are different, the result is strikingly similar. Python has built-in generators (which are used like Ruby blocks, as noted above) Python has support for generators in the language. In Ruby you could use the generator module that uses continuations to create a generator from a block. Or, you could just use a block/proc/lambda! Moreover, in Ruby 1.9 Fibers are, and can be used as, generators. docs.python.org has this generator example: def reverse(data): for index in range(len(data)-1, -1, -1): yield data[index] Contrast this with the above block examples. Python has flexible name space handling In Ruby, when you import a file with require, all the things defined in that file will end up in your global namespace. This causes namespace pollution. The solution to that is Rubys modules. But if you create a namespace with a module, then you have to use that namespace to access the contained classes. In Python, the file is a module, and you can import its contained names with from themodule import *, thereby polluting the namespace if you want. But you can also import just selected names with from themodule import aname, another or you can simply import themodule and then access the names with themodule.aname. If you want more levels in your namespace you can have packages, which are directories with modules and an __init__.py file. Python has docstrings Docstrings are strings that are attached to modules, functions and methods and can be introspected at runtime. This helps for creating such things as the help command and automatic documentation. def frobnicate(bar): """frobnicate takes a bar and frobnicates it >>> bar = Bar() >>> bar.is_frobnicated() False >>> frobnicate(bar) >>> bar.is_frobnicated() True """ Python has more libraries Python has a vast amount of available modules and bindings for libraries. Python has multiple inheritance Ruby does not ("on purpose" -- see Ruby's website, see here how it's done in Ruby). It does reuse the module concept as a sort of abstract classes. Python has list/dict comprehensions Python: res = [x*x for x in range(1, 10)] Ruby: res = (0..9).map { |x| x * x } Python: >>> (x*x for x in range(10)) <generator object <genexpr> at 0xb7c1ccd4> >>> list(_) [0, 1, 4, 9, 16, 25, 36, 49, 64, 81] Ruby: p = proc { |x| x * x } (0..9).map(&p) Python: >>> {x:str(y*y) for x,y in {1:2, 3:4}.items()} {1: '4', 3: '16'} Ruby: >> Hash[{1=>2, 3=>4}.map{|x,y| [x,(y*y).to_s]}] => {1=>"4", 3=>"16"} Python has decorators Things similar to decorators can be created in Ruby, and it can also be argued that they aren't as necessary as in Python.

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  • What statistics can be maintained for a set of numerical data without iterating?

    - by Dan Tao
    Update Just for future reference, I'm going to list all of the statistics that I'm aware of that can be maintained in a rolling collection, recalculated as an O(1) operation on every addition/removal (this is really how I should've worded the question from the beginning): Obvious Count Sum Mean Max* Min* Median** Less Obvious Variance Standard Deviation Skewness Kurtosis Mode*** Weighted Average Weighted Moving Average**** OK, so to put it more accurately: these are not "all" of the statistics I'm aware of. They're just the ones that I can remember off the top of my head right now. *Can be recalculated in O(1) for additions only, or for additions and removals if the collection is sorted (but in this case, insertion is not O(1)). Removals potentially incur an O(n) recalculation for non-sorted collections. **Recalculated in O(1) for a sorted, indexed collection only. ***Requires a fairly complex data structure to recalculate in O(1). ****This can certainly be achieved in O(1) for additions and removals when the weights are assigned in a linearly descending fashion. In other scenarios, I'm not sure. Original Question Say I maintain a collection of numerical data -- let's say, just a bunch of numbers. For this data, there are loads of calculated values that might be of interest; one example would be the sum. To get the sum of all this data, I could... Option 1: Iterate through the collection, adding all the values: double sum = 0.0; for (int i = 0; i < values.Count; i++) sum += values[i]; Option 2: Maintain the sum, eliminating the need to ever iterate over the collection just to find the sum: void Add(double value) { values.Add(value); sum += value; } void Remove(double value) { values.Remove(value); sum -= value; } EDIT: To put this question in more relatable terms, let's compare the two options above to a (sort of) real-world situation: Suppose I start listing numbers out loud and ask you to keep them in your head. I start by saying, "11, 16, 13, 12." If you've just been remembering the numbers themselves and nothing more, and then I say, "What's the sum?", you'd have to think to yourself, "OK, what's 11 + 16 + 13 + 12?" before responding, "52." If, on the other hand, you had been keeping track of the sum yourself while I was listing the numbers (i.e., when I said, "11" you thought "11", when I said "16", you thought, "27," and so on), you could answer "52" right away. Then if I say, "OK, now forget the number 16," if you've been keeping track of the sum inside your head you can simply take 16 away from 52 and know that the new sum is 36, rather than taking 16 off the list and them summing up 11 + 13 + 12. So my question is, what other calculations, other than the obvious ones like sum and average, are like this? SECOND EDIT: As an arbitrary example of a statistic that (I'm almost certain) does require iteration -- and therefore cannot be maintained as simply as a sum or average -- consider if I asked you, "how many numbers in this collection are divisible by the min?" Let's say the numbers are 5, 15, 19, 20, 21, 25, and 30. The min of this set is 5, which divides into 5, 15, 20, 25, and 30 (but not 19 or 21), so the answer is 5. Now if I remove 5 from the collection and ask the same question, the answer is now 2, since only 15 and 30 are divisible by the new min of 15; but, as far as I can tell, you cannot know this without going through the collection again. So I think this gets to the heart of my question: if we can divide kinds of statistics into these categories, those that are maintainable (my own term, maybe there's a more official one somewhere) versus those that require iteration to compute any time a collection is changed, what are all the maintainable ones? What I am asking about is not strictly the same as an online algorithm (though I sincerely thank those of you who introduced me to that concept). An online algorithm can begin its work without having even seen all of the input data; the maintainable statistics I am seeking will certainly have seen all the data, they just don't need to reiterate through it over and over again whenever it changes.

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  • Problem implementing Blinn–Phong shading model

    - by Joe Hopfgartner
    I did this very simple, perfectly working, implementation of Phong Relflection Model (There is no ambience implemented yet, but that doesn't bother me for now). The functions should be self explaining. /** * Implements the classic Phong illumination Model using a reflected light * vector. */ public class PhongIllumination implements IlluminationModel { @RGBParam(r = 0, g = 0, b = 0) public Vec3 ambient; @RGBParam(r = 1, g = 1, b = 1) public Vec3 diffuse; @RGBParam(r = 1, g = 1, b = 1) public Vec3 specular; @FloatParam(value = 20, min = 1, max = 200.0f) public float shininess; /* * Calculate the intensity of light reflected to the viewer . * * @param P = The surface position expressed in world coordinates. * * @param V = Normalized viewing vector from surface to eye in world * coordinates. * * @param N = Normalized normal vector at surface point in world * coordinates. * * @param surfaceColor = surfaceColor Color of the surface at the current * position. * * @param lights = The active light sources in the scene. * * @return Reflected light intensity I. */ public Vec3 shade(Vec3 P, Vec3 V, Vec3 N, Vec3 surfaceColor, Light lights[]) { Vec3 surfaceColordiffused = Vec3.mul(surfaceColor, diffuse); Vec3 totalintensity = new Vec3(0, 0, 0); for (int i = 0; i < lights.length; i++) { Vec3 L = lights[i].calcDirection(P); N = N.normalize(); V = V.normalize(); Vec3 R = Vec3.reflect(L, N); // reflection vector float diffuseLight = Vec3.dot(N, L); float specularLight = Vec3.dot(V, R); if (diffuseLight > 0) { totalintensity = Vec3.add(Vec3.mul(Vec3.mul( surfaceColordiffused, lights[i].calcIntensity(P)), diffuseLight), totalintensity); if (specularLight > 0) { Vec3 Il = lights[i].calcIntensity(P); Vec3 Ilincident = Vec3.mul(Il, Math.max(0.0f, Vec3 .dot(N, L))); Vec3 intensity = Vec3.mul(Vec3.mul(specular, Ilincident), (float) Math.pow(specularLight, shininess)); totalintensity = Vec3.add(totalintensity, intensity); } } } return totalintensity; } } Now i need to adapt it to become a Blinn-Phong illumination model I used the formulas from hearn and baker, followed pseudocodes and tried to implement it multiple times according to wikipedia articles in several languages but it never worked. I just get no specular reflections or they are so weak and/or are at the wrong place and/or have the wrong color. From the numerous wrong implementations I post some little code that already seems to be wrong. So I calculate my Half Way vector and my new specular light like so: Vec3 H = Vec3.mul(Vec3.add(L.normalize(), V), Vec3.add(L.normalize(), V).length()); float specularLight = Vec3.dot(H, N); With theese little changes it should already work (maby not with correct intensity but basically it should be correct). But the result is wrong. Here are two images. Left how it should render correctly and right how it renders. If i lower the shininess factor you can see a little specular light at the top right: Altough I understand the concept of Phong illumination and also the simplified more performant adaptaion of blinn phong I am trying around for days and just cant get it to work. Any help is appriciated. Edit: I was made aware of an error by this answer, that i am mutiplying by |L+V| instead of dividing by it when calculating H. I changed to deviding doing so: Vec3 H = Vec3.mul(Vec3.add(L.normalize(), V), 1/Vec3.add(L.normalize(), V).length()); Unfortunately this doesnt change much. The results look like this: and if I rise the specular constant and lower the shininess You can see the effects more clearly in a smilar wrong way: However this division just the normalisation. I think I am missing one step. Because the formulas like this just dont make sense to me. If you look at this picture: http://en.wikipedia.org/wiki/File:Blinn-Phong_vectors.svg The projection of H to N is far less than V to R. And if you imagine changing the vector V in the picture the angle is the same when the viewing vector is "on the left side". and becomes more and more different when going to the right. I pesonally would multiply the whole projection by two to become something similiar (and the hole point is to avoid the calculation of R). Altough I didnt read anythinga bout that anywehre i am gonna try this out... Result: The intension of the specular light is far too much (white areas) and the position is still wrong. I think I am messing something else up because teh reflection are just at the wrong place. But what? Edit: Now I read on wikipedia in the notes that the angle of N/H is in fact approximalty half or V/R. To compensate that i should multiply my shineness exponent by 4 rather than my projection. If i do that I end up with this: Far to intense but still one thing. The projection is at the wrong place. Where could i mess up my vectors?

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  • How to filter Backbone.js Collection and Rerender App View?

    - by Jeremy H.
    Is is a total Backbone.js noob question. I am working off of the ToDo Backbone.js example trying to build out a fairly simple single app interface. While the todo project is more about user input, this app is more about filtering the data based on the user options (click events). I am completely new to Backbone.js and Mongoose and have been unable to find a good example of what I am trying to do. I have been able to get my api to pull the data from the MongoDB collection and drop it into the Backbone.js collection which renders it in the app. What I cannot for the life of me figure out how to do is filter that data and re-render the app view. I am trying to filter by the "type" field in the document. Here is my script: (I am totally aware of some major refactoring needed, I am just rapid prototyping a concept.) $(function() { window.Job = Backbone.Model.extend({ idAttribute: "_id", defaults: function() { return { attachments: false } } }); window.JobsList = Backbone.Collection.extend({ model: Job, url: '/api/jobs', leads: function() { return this.filter(function(job){ return job.get('type') == "Lead"; }); } }); window.Jobs = new JobsList; window.JobView = Backbone.View.extend({ tagName: "div", className: "item", template: _.template($('#item-template').html()), initialize: function() { this.model.bind('change', this.render, this); this.model.bind('destroy', this.remove, this); }, render: function() { $(this.el).html(this.template(this.model.toJSON())); this.setText(); return this; }, setText: function() { var month=new Array(); month[0]="Jan"; month[1]="Feb"; month[2]="Mar"; month[3]="Apr"; month[4]="May"; month[5]="Jun"; month[6]="Jul"; month[7]="Aug"; month[8]="Sep"; month[9]="Oct"; month[10]="Nov"; month[11]="Dec"; var title = this.model.get('title'); var description = this.model.get('description'); var datemonth = this.model.get('datem'); var dateday = this.model.get('dated'); var jobtype = this.model.get('type'); var jobstatus = this.model.get('status'); var amount = this.model.get('amount'); var paymentstatus = this.model.get('paymentstatus') var type = this.$('.status .jobtype'); var status = this.$('.status .jobstatus'); this.$('.title a').text(title); this.$('.description').text(description); this.$('.date .month').text(month[datemonth]); this.$('.date .day').text(dateday); type.text(jobtype); status.text(jobstatus); if(amount > 0) this.$('.paymentamount').text(amount) if(paymentstatus) this.$('.paymentstatus').text(paymentstatus) if(jobstatus === 'New') { status.addClass('new'); } else if (jobstatus === 'Past Due') { status.addClass('pastdue') }; if(jobtype === 'Lead') { type.addClass('lead'); } else if (jobtype === '') { type.addClass(''); }; }, remove: function() { $(this.el).remove(); }, clear: function() { this.model.destroy(); } }); window.AppView = Backbone.View.extend({ el: $("#main"), events: { "click #leads .highlight" : "filterLeads" }, initialize: function() { Jobs.bind('add', this.addOne, this); Jobs.bind('reset', this.addAll, this); Jobs.bind('all', this.render, this); Jobs.fetch(); }, addOne: function(job) { var view = new JobView({model: job}); this.$("#activitystream").append(view.render().el); }, addAll: function() { Jobs.each(this.addOne); }, filterLeads: function() { // left here, this event fires but i need to figure out how to filter the activity list. } }); window.App = new AppView; });

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  • Trying to reduce the speed overhead of an almost-but-not-quite-int number class

    - by Fumiyo Eda
    I have implemented a C++ class which behaves very similarly to the standard int type. The difference is that it has an additional concept of "epsilon" which represents some tiny value that is much less than 1, but greater than 0. One way to think of it is as a very wide fixed point number with 32 MSBs (the integer parts), 32 LSBs (the epsilon parts) and a huge sea of zeros in between. The following class works, but introduces a ~2x speed penalty in the overall program. (The program includes code that has nothing to do with this class, so the actual speed penalty of this class is probably much greater than 2x.) I can't paste the code that is using this class, but I can say the following: +, -, +=, <, > and >= are the only heavily used operators. Use of setEpsilon() and getInt() is extremely rare. * is also rare, and does not even need to consider the epsilon values at all. Here is the class: #include <limits> struct int32Uepsilon { typedef int32Uepsilon Self; int32Uepsilon () { _value = 0; _eps = 0; } int32Uepsilon (const int &i) { _value = i; _eps = 0; } void setEpsilon() { _eps = 1; } Self operator+(const Self &rhs) const { Self result = *this; result._value += rhs._value; result._eps += rhs._eps; return result; } Self operator-(const Self &rhs) const { Self result = *this; result._value -= rhs._value; result._eps -= rhs._eps; return result; } Self operator-( ) const { Self result = *this; result._value = -result._value; result._eps = -result._eps; return result; } Self operator*(const Self &rhs) const { return this->getInt() * rhs.getInt(); } // XXX: discards epsilon bool operator<(const Self &rhs) const { return (_value < rhs._value) || (_value == rhs._value && _eps < rhs._eps); } bool operator>(const Self &rhs) const { return (_value > rhs._value) || (_value == rhs._value && _eps > rhs._eps); } bool operator>=(const Self &rhs) const { return (_value >= rhs._value) || (_value == rhs._value && _eps >= rhs._eps); } Self &operator+=(const Self &rhs) { this->_value += rhs._value; this->_eps += rhs._eps; return *this; } Self &operator-=(const Self &rhs) { this->_value -= rhs._value; this->_eps -= rhs._eps; return *this; } int getInt() const { return(_value); } private: int _value; int _eps; }; namespace std { template<> struct numeric_limits<int32Uepsilon> { static const bool is_signed = true; static int max() { return 2147483647; } } }; The code above works, but it is quite slow. Does anyone have any ideas on how to improve performance? There are a few hints/details I can give that might be helpful: 32 bits are definitely insufficient to hold both _value and _eps. In practice, up to 24 ~ 28 bits of _value are used and up to 20 bits of _eps are used. I could not measure a significant performance difference between using int32_t and int64_t, so memory overhead itself is probably not the problem here. Saturating addition/subtraction on _eps would be cool, but isn't really necessary. Note that the signs of _value and _eps are not necessarily the same! This broke my first attempt at speeding this class up. Inline assembly is no problem, so long as it works with GCC on a Core i7 system running Linux!

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  • Gradient algororithm produces little white dots

    - by user146780
    I'm working on an algorithm to generate point to point linear gradients. I have a rough, proof of concept implementation done: GLuint OGLENGINEFUNCTIONS::CreateGradient( std::vector<ARGBCOLORF> &input,POINTFLOAT start, POINTFLOAT end, int width, int height,bool radial ) { std::vector<POINT> pol; std::vector<GLubyte> pdata(width * height * 4); std::vector<POINTFLOAT> linearpts; std::vector<float> lookup; float distance = GetDistance(start,end); RoundNumber(distance); POINTFLOAT temp; float incr = 1 / (distance + 1); for(int l = 0; l < 100; l ++) { POINTFLOAT outA; POINTFLOAT OutB; float dirlen; float perplen; POINTFLOAT dir; POINTFLOAT ndir; POINTFLOAT perp; POINTFLOAT nperp; POINTFLOAT perpoffset; POINTFLOAT diroffset; dir.x = end.x - start.x; dir.y = end.y - start.y; dirlen = sqrt((dir.x * dir.x) + (dir.y * dir.y)); ndir.x = static_cast<float>(dir.x * 1.0 / dirlen); ndir.y = static_cast<float>(dir.y * 1.0 / dirlen); perp.x = dir.y; perp.y = -dir.x; perplen = sqrt((perp.x * perp.x) + (perp.y * perp.y)); nperp.x = static_cast<float>(perp.x * 1.0 / perplen); nperp.y = static_cast<float>(perp.y * 1.0 / perplen); perpoffset.x = static_cast<float>(nperp.x * l * 0.5); perpoffset.y = static_cast<float>(nperp.y * l * 0.5); diroffset.x = static_cast<float>(ndir.x * 0 * 0.5); diroffset.y = static_cast<float>(ndir.y * 0 * 0.5); outA.x = end.x + perpoffset.x + diroffset.x; outA.y = end.y + perpoffset.y + diroffset.y; OutB.x = start.x + perpoffset.x - diroffset.x; OutB.y = start.y + perpoffset.y - diroffset.y; for (float i = 0; i < 1; i += incr) { temp = GetLinearBezier(i,outA,OutB); RoundNumber(temp.x); RoundNumber(temp.y); linearpts.push_back(temp); lookup.push_back(i); } for (unsigned int j = 0; j < linearpts.size(); j++) { if(linearpts[j].x < width && linearpts[j].x >= 0 && linearpts[j].y < height && linearpts[j].y >=0) { pdata[linearpts[j].x * 4 * width + linearpts[j].y * 4 + 0] = (GLubyte) j; pdata[linearpts[j].x * 4 * width + linearpts[j].y * 4 + 1] = (GLubyte) j; pdata[linearpts[j].x * 4 * width + linearpts[j].y * 4 + 2] = (GLubyte) j; pdata[linearpts[j].x * 4 * width + linearpts[j].y * 4 + 3] = (GLubyte) 255; } } lookup.clear(); linearpts.clear(); } return CreateTexture(pdata,width,height); } It works as I would expect most of the time, but at certain angles it produces little white dots. I can't figure out what does this. This is what it looks like at most angles (good) http://img9.imageshack.us/img9/5922/goodgradient.png But once in a while it looks like this (bad): http://img155.imageshack.us/img155/760/badgradient.png What could be causing the white dots? Is there maybe also a better way to generate my gradients if no solution is possible for this? Thanks

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  • Replacing objects, handling clones, dealing with write logs

    - by Alix
    Hi everyone, I'm dealing with a problem I can't figure out how to solve, and I'd love to hear some suggestions. [NOTE: I realise I'm asking several questions; however, answers need to take into account all of the issues, so I cannot split this into several questions] Here's the deal: I'm implementing a system that underlies user applications and that protect shared objects from concurrent accesses. The application programmer (whose application will run on top of my system) defines such shared objects like this: public class MyAtomicObject { // These are just examples of fields you may want to have in your class. public virtual int x { get; set; } public virtual List<int> list { get; set; } public virtual MyClassA objA { get; set; } public virtual MyClassB objB { get; set; } } As you can see they declare the fields of their class as auto-generated properties (auto-generated means they don't need to implement get and set). This is so that I can go in and extend their class and implement each get and set myself in order to handle possible concurrent accesses, etc. This is all well and good, but now it starts to get ugly: the application threads run transactions, like this: The thread signals it's starting a transaction. This means we now need to monitor its accesses to the fields of the atomic objects. The thread runs its code, possibly accessing fields for reading or writing. If there are accesses for writing, we'll hide them from the other transactions (other threads), and only make them visible in step 3. This is because the transaction may fail and have to roll back (undo) its updates, and in that case we don't want other threads to see its "dirty" data. The thread signals it wants to commit the transaction. If the commit is successful, the updates it made will now become visible to everyone else. Otherwise, the transaction will abort, the updates will remain invisible, and no one will ever know the transaction was there. So basically the concept of transaction is a series of accesses that appear to have happened atomically, that is, all at the same time, in the same instant, which would be the moment of successful commit. (This is as opposed to its updates becoming visible as it makes them) In order to hide the write accesses in step 2, I clone the accessed field (let's say it's the field list) and put it in the transaction's write log. After that, any time the transaction accesses list, it will actually be accessing the clone in its write log, and not the global copy everyone else sees. Like this, any changes it makes will be done to the (invisible) clone, not to the global copy. If in step 3 the commit is successful, the transaction should replace the global copy with the updated list it has in its write log, and then the changes become visible for everyone else at once. It would be something like this: myAtomicObject.list = updatedCloneOfListInTheWriteLog; Problem #1: possible references to the list. Let's say someone puts a reference to the global list in a dictionary. When I do... myAtomicObject.list = updatedCloneOfListInTheWriteLog; ...I'm just replacing the reference in the field list, but not the real object (I'm not overwriting the data), so in the dictionary we'll still have a reference to the old version of the list. A possible solution would be to overwrite the data (in the case of a list, empty the global list and add all the elements of the clone). More generically, I would need to copy the fields of one list to the other. I can do this with reflection, but that's not very pretty. Is there any other way to do it? Problem #2: even if problem #1 is solved, I still have a similar problem with the clone: the application programmer doesn't know I'm giving him a clone and not the global copy. What if he puts the clone in a dictionary? Then at commit there will be some references to the global copy and some to the clone, when in truth they should all point to the same object. I thought about providing a wrapper object that contains both the cloned list and a pointer to the global copy, but the programmer doesn't know about this wrapper, so they're not going to use the pointer at all. The wrapper would be like this: public class Wrapper<T> : T { // This would be the pointer to the global copy. The local data is contained in whatever fields the wrapper inherits from T. private T thisPtr; } I do need this wrapper for comparisons: if I have a dictionary that has an entry with the global copy as key, if I look it up with the clone, like this: dictionary[updatedCloneOfListInTheWriteLog] I need it to return the entry, that is, to think that updatedCloneOfListInTheWriteLog and the global copy are the same thing. For this, I can just override Equals, GetHashCode, operator== and operator!=, no problem. However I still don't know how to solve the case in which the programmer unknowingly inserts a reference to the clone in a dictionary. Problem #3: the wrapper must extend the class of the object it wraps (if it's wrapping MyClassA, it must extend MyClassA) so that it's accepted wherever an object of that class (MyClass) would be accepted. However, that class (MyClassA) may be final. This is pretty horrible :$. Any suggestions? I don't need to use a wrapper, anything you can think of is fine. What I cannot change is the write log (I need to have a write log) and the fact that the programmer doesn't know about the clone. I hope I've made some sense. Feel free to ask for more info if something needs some clearing up. Thanks so much!

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  • Google App Engine - SiteMap Creation for a social network

    - by spidee
    Hi all. I am creating a social tool - I want to allow search engines to pick up "public" user profiles - like twitter and face-book. I have seen all the protocol info at http://www.sitemaps.org and i understand this and how to build such a file - along with an index if i exceed the 50K limit. Where i am struggling is the concept of how i make this run. The site map for my general site pages is simple i can use a tool to create the file - or a script - host the file - submit the file and done. What i then need is a script that will create the site-maps of user profiles. I assume this would be something like: <?xml version="1.0" encoding="UTF-8"?> <urlset xmlns="http://www.sitemaps.org/schemas/sitemap/0.9"> <url> <loc>http://www.socialsite.com/profile/spidee</loc> <lastmod>2010-5-12</lastmod> <changefreq>???</changefreq> <priority>???</priority> </url> <url> <loc>http://www.socialsite.com/profile/webbsterisback</loc> <lastmod>2010-5-12</lastmod> <changefreq>???</changefreq> <priority>???</priority> </url> </urlset> Ive added some ??? as i don't know how i should set these settings for my profiles based on the following:- When a new profile is created it must be added to a site-map. If the profile is changed or if "certain" properties are changed - then i don't know if i update the entry in the map - or do something else? (updating would be a nightmare!) Some users may change their profile. In terms of relevance to the search engine the only way a google or yahoo search will find the users (for my requirement) profile would be for example by means of [user name] and [location] so once the entry for the profile has been added to the map file the only reason to have the search-bot re-index the profile would be if the user changed their user-name - which they cant. or their location - and or set their settings so that their profile would be "hidden" from search engines. I assume my map creation will need to be dynamic. From what i have said above i would imagine that creating a new profile and possible editing certain properties could mark it as needing adding/updating in the sitemap. Assuming i will have millions of profiles added/being edited how can i manage this in a sensible manner. i know i need a script that can append urls as each profile is created i know the script will prob be a TASK - running at a set freq - perhaps the profiles have a property like "indexed" and the TASK sets them to "true" when the profiles are added to the map. I dont see the best way to store the map - do i store it in the datastore i.e; model=sitemaps properties key_name=sitemap_xml_1 (and for my map sitemap_index_xml) mapxml=blobstore (the raw xml map or ror map) full=boolean (set true when url count is 50) # might need this as a shard will tell us To make this work my thoughts are m cache the current site map structure as "sitemap_xml" keep a shard of url count when my task executes 1. build the xml structure for say the first 100 urls marked "index==false" (how many could u run at a time?) 2. test if the current mcache sitemap is full (shardcounter+10050K) 3.a if the map is near full create a new map entry in models "sitemap_xml_2" - update the map_index file (also stored in my model as "sitemap_index" start a new shard - or reset.2 3.b if the map is not full grab it from mcache 4.append the 100 url xml structure 5.save / m cache the map I can now add a handler using a url map/route like /sitemaps/* Get my * as map name and serve the maps from the blobstore/mache on the fly. Now my question is does this work - is this the right way or a good way to start? Will this handle the situation of making sure the search bots update when a user changes their profile - possibly by setting the change freq correctly? - Do i need a more advance system :( ? or have i re-invented the wheel! I hope this is all clear and make some form of sense :-)

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  • How to add a 3rd level to my CSS drop down menu?

    - by Cynthia
    I have a 2-level drop down menu that looks great in all browsers. Now I want to add a 3rd level. How do I do that? Here is my HTML for the menu: <div class="nav"> <div class="navbar"> <ul class="menu"> <li><a href="#">Home</a></li> <li><a href="#">About JoyFactory</a> <ul class="sub-menu"> <li><a href="#">Who We Are</a></li> <li><a href="#">Our Education Concept</a></li> <li><a href="#">References</a></li> </ul> </li> <li><a href="#">JoyFactory Kinderkrippe</a> <ul class="sub-menu"> <li><a href="#">JoyFactory Kinderkrippe Oerlikon</a> <ul> <li><a href="#">item 1</a></li> <li><a href="#">item 2</a></li> <li><a href="#">item 3</a></li> <li><a href="#">item 4</a></li> </ul> </li> <li><a href="#">JoyFactory Kinderkrippe Seebach</a></li> </ul> </li> </ul> </div> </div> and here is my CSS: .nav { clear:both ; width:1020px ; height:55px ; background:url("images/nav-bg.png") no-repeat ; position:absolute ; top:125px ; left:-10px ; } .navbar { width:1000px ; height:50px ; margin:auto ; } ul.menu { margin-left:0 ; padding-left:0 ; list-style-type:none ; } .menu li { display:inline ; float:left ; height:50px ; margin:0 6px ; } .menu li a { font-family:'MyriadPro-SemiboldCond' ; font-size:24px ; color:#ffffff ; text-decoration:none ; height:50px ; line-height:50px ; padding:0px 10px ; } .menu li:hover, .menu li:hover a { background:#ffd322 ; color:#e32a0e ; } .sub-menu { position:absolute ; float:none ; padding:0 ; top:50px ; z-index:9999 ; background:#ffd322 ; margin-left:0 ; padding-left:0 ; } .sub-menu li { display:none ; min-width:175px !important ; margin: 0 !important; padding: 0 !important; } .sub-menu li a, .current-menu-parent .sub-menu li a { display:block ; background:#ffd322 ; font-family:arial,helvetica,sans-serif ; font-size:16px ; padding:0 10px ; border-top:1px solid #f37f10 ; border-left:none ; } .sub-menu li a:hover, .menu li.current-menu-parent .sub-menu li.current-menu-item a { background:#f37f10 } .menu li:hover li { float: none; display:block; clear: both; } Any help would be most appreciated! Many thanks :)

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  • Does this language feature already exist?

    - by Pindatjuh
    I'm currently developing a new language for programming in a continuous environment (compare it to electrical engineering), and I've got some ideas on a certain language construction. Let me explain the feature by explanation and then by definition: x = a U b; Where x is a variable and a and b are other variables (or static values). This works like a union between a and b; no duplicates and no specific order. with(x) { // regular 'with' usage; using the global interpretation of "x" x = 5; // will replace the original definition of "x = a U b;" } with(x = a) { // this code block is executed when the "x" variable // has the "a" variable assigned. All references in // this code-block to "x" are references to "a". So saying: x = 5; // would only change the variable "a". If the variable "a" // later on changes, x still equals to 5, in this fashion: // 'x = a U b U 5;' // '[currentscope] = 5;' // thus, 'a = 5;' } with(x = b) { // same but with "b" } with(x != a) { // here the "x" variable refers to any variable // but "a"; thus saying x = 5; // is equal to the rewriting of // 'x = a U b U 5;' // 'b = 5;' (since it was the scope of this block) } with(x = (a U b)) { // guaranteed that "x" is 'a U b'; interacting with "x" // will interact with both "a" and "b". x = 5; // makes both "a" and "b" equal to 5; also the "x" variable // is updated to contain: // 'x = a U b U 5;' // '[currentscope] = 5;' // 'a U b = 5;' // and thus: 'a = 5; b = 5;'. } // etc. In the above, all code-blocks are executed, but the "scope" changes in each block how x is interpreted. In the first block, x is guaranteed to be a: thus interacting with x inside that block will interact on a. The second and the third code-block are only equal in this situation (because not a: then there only remains b). The last block guarantees that x is at least a or b. Further more; U is not the "bitwise or operator", but I've called it the "and/or"-operator. Its definition is: "U" = "and" U "or" (On my blog, http://cplang.wordpress.com/2009/12/19/binop-and-or/, there is more (mathematical) background information on this operator. It's loosely based on sets. Using different syntax, changed it in this question.) Update: more examples. print = "Hello world!" U "How are you?"; // this will print // both values, but the // order doesn't matter. // 'userkey' is a variable containing a key. with(userkey = "a") { print = userkey; // will only print "a". } with(userkey = ("shift" U "a")) { // pressed both "shift" and the "a" key. print = userkey; // will "print" shift and "a", even // if the user also pressed "ctrl": // the interpretation of "userkey" is changed, // such that it only contains the matched cases. } with((userkey = "shift") U (userkey = "a")) { // same as if-statement above this one, showing the distributivity. } x = 5 U 6 U 7; y = x + x; // will be: // y = (5 U 6 U 7) + (5 U 6 U 7) // = 10 U 11 U 12 U 13 U 14 somewantedkey = "ctrl" U "alt" U "space" with(userkey = somewantedkey) { // must match all elements of "somewantedkey" // (distributed the Boolean equals operated) // thus only executed when all the defined keys are pressed } with(somewantedkey = userkey) { // matches only one of the provided "somewantedkey" // thus when only "space" is pressed, this block is executed. } Update2: more examples and some more context. with(x = (a U b)) { // this } // can be written as with((x = a) U (x = b)) { // this: changing the variable like x = 5; // will be rewritten as: // a = 5 and b = 5 } Some background information: I'm building a language which is "time-independent", like Java is "platform-independant". Everything stated in the language is "as is", and is continuously actively executed. This means; the programmer does not know in which order (unless explicitly stated using constructions) elements are, nor when statements are executed. The language is completely separated from the "time"-concept, i.e. it's continuously executed: with(a < 5) { a++; } // this is a loop-structure; // how and when it's executed isn't known however. with(a) { // everytime the "a" variable changes, this code-block is executed. b = 4; with(b < 3) { // runs only three times. } with(b > 0) { b = b - 1; // runs four times } } Update 3: After pondering on the type of this language feature; it closely resemblances Netbeans Platform's Lookup, where each "with"-statement a synchronized agent is, working on it's specific "filter" of objects. Instead of type-based, this is variable-based (fundamentally quite the same; just a different way of identifiying objects). I greatly thank all of you for providing me with very insightful information and links/hints to great topics I can research. Thanks. I do not know if this construction already exists, so that's my question: does this language feature already exist?

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  • setting up linked list Java

    - by erp
    I'm working on some basic linked list stuff, like insert, delete, go to the front or end of the list, and basically i understand the concept of all of that stuff once i have the list i guess but im having trouble setting up the list. I was wondering of you guys could tell me if im going in the right direction. (mostly just the setup) this is what i have so far: public class List { private int size; private List linkedList; List head; List cur; List next; /** * Creates an empty list. * @pre * @post */ public List(){ linkedList = new List(); this.head = null; cur = head; } /** * Delete the current element from this list. The element after the deleted element becomes the new current. * If that's not possible, then the element before the deleted element becomes the new current. * If that is also not possible, then you need to recognize what state the list is in and define current accordingly. * Nothing should be done if a delete is not possible. * @pre * @post */ public void delete(){ // delete size--; } /** * Get the value of the current element. If this is not possible, throw an IllegalArgumentException. * @pre the list is not empty * @post * @return value of the current element. */ public char get(){ return getItem(cur); } /** * Go to the last element of the list. If this is not possible, don't change the cursor. * @pre * @post */ public void goLast(){ while (cur.next != null){ cur = cur.next; } } /** * Advance the cursor to the next element. If this is not possible, don't change the cursor. * @pre * @post */ public void goNext(){ if(cur.next != null){ cur = cur.next;} //else do nothing } /** * Retreat the cursor to the previous element. If this is not possible, don't change the cursor. * @pre * @post */ public void goPrev(){ } /** * Go to top of the list. This is the position before the first element. * @pre * @post */ public void goTop(){ } /** * Go to first element of the list. If this is not possible, don't change the cursor. * @pre * @post */ public void goFirst(){ } /** * Insert the given parameter after the current element. The newly inserted element becomes the current element. * @pre * @post * @param newVal : value to insert after the current element. */ public void insert(char newVal){ cur.setItem(newVal); size++; } /** * Determines if this list is empty. Empty means this list has no elements. * @pre * @post * @return true if the list is empty. */ public boolean isEmpty(){ return head == null; } /** * Determines the size of the list. The size of the list is the number of elements in the list. * @pre * @post * @return size which is the number of elements in the list. */ public int size(){ return size; } public class Node { private char item; private Node next; public Node() { } public Node(char item) { this.item = item; } public Node(char item, Node next) { this.item = item; this.next = next; } public char getItem() { return this.item; } public void setItem(char item) { this.item = item; } public Node getNext() { return this.next; } public void setNext(Node next) { this.next = next; } } } I got the node class alright (well i think it works alright), but is it necessary to even have that class? or can i go about it without even using it (just curious). And for example on the method get() in the list class can i not call that getItem() method from the node class because it's getting an error even though i thought that was the whole point for the node class. bottom line i just wanna make sure im setting up the list right. Thanks for any help guys, im new to linked list's so bear with me!

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  • Loading jQuery Consistently in a .NET Web App

    - by Rick Strahl
    One thing that frequently comes up in discussions when using jQuery is how to best load the jQuery library (as well as other commonly used and updated libraries) in a Web application. Specifically the issue is the one of versioning and making sure that you can easily update and switch versions of script files with application wide settings in one place and having your script usage reflect those settings in the entire application on all pages that use the script. Although I use jQuery as an example here, the same concepts can be applied to any script library - for example in my Web libraries I use the same approach for jQuery.ui and my own internal jQuery support library. The concepts used here can be applied both in WebForms and MVC. Loading jQuery Properly From CDN Before we look at a generic way to load jQuery via some server logic, let me first point out my preferred way to embed jQuery into the page. I use the Google CDN to load jQuery and then use a fallback URL to handle the offline or no Internet connection scenario. Why use a CDN? CDN links tend to be loaded more quickly since they are very likely to be cached in user's browsers already as jQuery CDN is used by many, many sites on the Web. Using a CDN also removes load from your Web server and puts the load bearing on the CDN provider - in this case Google - rather than on your Web site. On the downside, CDN links gives the provider (Google, Microsoft) yet another way to track users through their Web usage. Here's how I use jQuery CDN plus a fallback link on my WebLog for example: <!DOCTYPE HTML> <html> <head> <script src="//ajax.googleapis.com/ajax/libs/jquery/1.6.4/jquery.min.js"></script> <script> if (typeof (jQuery) == 'undefined') document.write(unescape("%3Cscript " + "src='/Weblog/wwSC.axd?r=Westwind.Web.Controls.Resources.jquery.js' %3E%3C/script%3E")); </script> <title>Rick Strahl's Web Log</title> ... </head>   You can see that the CDN is referenced first, followed by a small script block that checks to see whether jQuery was loaded (jQuery object exists). If it didn't load another script reference is added to the document dynamically pointing to a backup URL. In this case my backup URL points at a WebResource in my Westwind.Web  assembly, but the URL can also be local script like src="/scripts/jquery.min.js". Important: Use the proper Protocol/Scheme for  for CDN Urls [updated based on comments] If you're using a CDN to load an external script resource you should always make sure that the script is loaded with the same protocol as the parent page to avoid mixed content warnings by the browser. You don't want to load a script link to an http:// resource when you're on an https:// page. The easiest way to use this is by using a protocol relative URL: <script src="//ajax.googleapis.com/ajax/libs/jquery/1.6.4/jquery.min.js"></script> which is an easy way to load resources from other domains. This URL syntax will automatically use the parent page's protocol (or more correctly scheme). As long as the remote domains support both http:// and https:// access this should work. BTW this also works in CSS (with some limitations) and links. BTW, I didn't know about this until it was pointed out in the comments. This is a very useful feature for many things - ah the benefits of my blog to myself :-) Version Numbers When you use a CDN you notice that you have to reference a specific version of jQuery. When using local files you may not have to do this as you can rename your private copy of jQuery.js, but for CDN the references are always versioned. The version number is of course very important to ensure you getting the version you have tested with, but it's also important to the provider because it ensures that cached content is always correct. If an existing file was updated the updates might take a very long time to get past the locally cached content and won't refresh properly. The version number ensures you get the right version and not some cached content that has been changed but not updated in your cache. On the other hand version numbers also mean that once you decide to use a new version of the script you now have to change all your script references in your pages. Depending on whether you use some sort of master/layout page or not this may or may not be easy in your application. Even if you do use master/layout pages, chances are that you probably have a few of them and at the very least all of those have to be updated for the scripts. If you use individual pages for all content this issue then spreads to all of your pages. Search and Replace in Files will do the trick, but it's still something that's easy to forget and worry about. Personaly I think it makes sense to have a single place where you can specify common script libraries that you want to load and more importantly which versions thereof and where they are loaded from. Loading Scripts via Server Code Script loading has always been important to me and as long as I can remember I've always built some custom script loading routines into my Web frameworks. WebForms makes this fairly easy because it has a reasonably useful script manager (ClientScriptManager and the ScriptManager) which allow injecting script into the page easily from anywhere in the Page cycle. What's nice about these components is that they allow scripts to be injected by controls so components can wrap up complex script/resource dependencies more easily without having to require long lists of CSS/Scripts/Image includes. In MVC or pure script driven applications like Razor WebPages  the process is more raw, requiring you to embed script references in the right place. But its also more immediate - it lets you know exactly which versions of scripts to use because you have to manually embed them. In WebForms with different controls loading resources this often can get confusing because it's quite possible to load multiple versions of the same script library into a page, the results of which are less than optimal… In this post I look a simple routine that embeds jQuery into the page based on a few application wide configuration settings. It returns only a string of the script tags that can be manually embedded into a Page template. It's a small function that merely a string of the script tags shown at the begging of this post along with some options on how that string is comprised. You'll be able to specify in one place which version loads and then all places where the help function is used will automatically reflect this selection. Options allow specification of the jQuery CDN Url, the fallback Url and where jQuery should be loaded from (script folder, Resource or CDN in my case). While this is specific to jQuery you can apply this to other resources as well. For example I use a similar approach with jQuery.ui as well using practically the same semantics. Providing Resources in ControlResources In my Westwind.Web Web utility library I have a class called ControlResources which is responsible for holding resource Urls, resource IDs and string contants that reference those resource IDs. The library also provides a few helper methods for loading common scriptscripts into a Web page. There are specific versions for WebForms which use the ClientScriptManager/ScriptManager and script link methods that can be used in any .NET technology that can embed an expression into the output template (or code for that matter). The ControlResources class contains mostly static content - references to resources mostly. But it also contains a few static properties that configure script loading: A Script LoadMode (CDN, Resource, or script url) A default CDN Url A fallback url They are  static properties in the ControlResources class: public class ControlResources { /// <summary> /// Determines what location jQuery is loaded from /// </summary> public static JQueryLoadModes jQueryLoadMode = JQueryLoadModes.ContentDeliveryNetwork; /// <summary> /// jQuery CDN Url on Google /// </summary> public static string jQueryCdnUrl = "//ajax.googleapis.com/ajax/libs/jquery/1.6.4/jquery.min.js"; /// <summary> /// jQuery CDN Url on Google /// </summary> public static string jQueryUiCdnUrl = "//ajax.googleapis.com/ajax/libs/jqueryui/1.8.16/jquery-ui.min.js"; /// <summary> /// jQuery UI fallback Url if CDN is unavailable or WebResource is used /// Note: The file needs to exist and hold the minimized version of jQuery ui /// </summary> public static string jQueryUiLocalFallbackUrl = "~/scripts/jquery-ui.min.js"; } These static properties are fixed values that can be changed at application startup to reflect your preferences. Since they're static they are application wide settings and respected across the entire Web application running. It's best to set these default in Application_Init or similar startup code if you need to change them for your application: protected void Application_Start(object sender, EventArgs e) { // Force jQuery to be loaded off Google Content Network ControlResources.jQueryLoadMode = JQueryLoadModes.ContentDeliveryNetwork; // Allow overriding of the Cdn url ControlResources.jQueryCdnUrl = "http://ajax.googleapis.com/ajax/libs/jquery/1.6.2/jquery.min.js"; // Route to our own internal handler App.OnApplicationStart(); } With these basic settings in place you can then embed expressions into a page easily. In WebForms use: <!DOCTYPE html> <html> <head runat="server"> <%= ControlResources.jQueryLink() %> <script src="scripts/ww.jquery.min.js"></script> </head> In Razor use: <!DOCTYPE html> <html> <head> @Html.Raw(ControlResources.jQueryLink()) <script src="scripts/ww.jquery.min.js"></script> </head> Note that in Razor you need to use @Html.Raw() to force the string NOT to escape. Razor by default escapes string results and this ensures that the HTML content is properly expanded as raw HTML text. Both the WebForms and Razor output produce: <!DOCTYPE html> <html> <head> <script src="http://ajax.googleapis.com/ajax/libs/jquery/1.6.2/jquery.min.js" type="text/javascript"></script> <script type="text/javascript"> if (typeof (jQuery) == 'undefined') document.write(unescape("%3Cscript src='/WestWindWebToolkitWeb/WebResource.axd?d=-b6oWzgbpGb8uTaHDrCMv59VSmGhilZP5_T_B8anpGx7X-PmW_1eu1KoHDvox-XHqA1EEb-Tl2YAP3bBeebGN65tv-7-yAimtG4ZnoWH633pExpJor8Qp1aKbk-KQWSoNfRC7rQJHXVP4tC0reYzVw2&t=634535391996872492' type='text/javascript'%3E%3C/script%3E"));</script> <script src="scripts/ww.jquery.min.js"></script> </head> which produces the desired effect for both CDN load and fallback URL. The implementation of jQueryLink is pretty basic of course: /// <summary> /// Inserts a script link to load jQuery into the page based on the jQueryLoadModes settings /// of this class. Default load is by CDN plus WebResource fallback /// </summary> /// <param name="url"> /// An optional explicit URL to load jQuery from. Url is resolved. /// When specified no fallback is applied /// </param> /// <returns>full script tag and fallback script for jQuery to load</returns> public static string jQueryLink(JQueryLoadModes jQueryLoadMode = JQueryLoadModes.Default, string url = null) { string jQueryUrl = string.Empty; string fallbackScript = string.Empty; if (jQueryLoadMode == JQueryLoadModes.Default) jQueryLoadMode = ControlResources.jQueryLoadMode; if (!string.IsNullOrEmpty(url)) jQueryUrl = WebUtils.ResolveUrl(url); else if (jQueryLoadMode == JQueryLoadModes.WebResource) { Page page = new Page(); jQueryUrl = page.ClientScript.GetWebResourceUrl(typeof(ControlResources), ControlResources.JQUERY_SCRIPT_RESOURCE); } else if (jQueryLoadMode == JQueryLoadModes.ContentDeliveryNetwork) { jQueryUrl = ControlResources.jQueryCdnUrl; if (!string.IsNullOrEmpty(jQueryCdnUrl)) { // check if jquery loaded - if it didn't we're not online and use WebResource fallbackScript = @"<script type=""text/javascript"">if (typeof(jQuery) == 'undefined') document.write(unescape(""%3Cscript src='{0}' type='text/javascript'%3E%3C/script%3E""));</script>"; fallbackScript = string.Format(fallbackScript, WebUtils.ResolveUrl(ControlResources.jQueryCdnFallbackUrl)); } } string output = "<script src=\"" + jQueryUrl + "\" type=\"text/javascript\"></script>"; // add in the CDN fallback script code if (!string.IsNullOrEmpty(fallbackScript)) output += "\r\n" + fallbackScript + "\r\n"; return output; } There's one dependency here on WebUtils.ResolveUrl() which resolves Urls without access to a Page/Control (another one of those features that should be in the runtime, not in the WebForms or MVC engine). You can see there's only a little bit of logic in this code that deals with potentially different load modes. I can load scripts from a Url, WebResources or - my preferred way - from CDN. Based on the static settings the scripts to embed are composed to be returned as simple string <script> tag(s). I find this extremely useful especially when I'm not connected to the internet so that I can quickly swap in a local jQuery resource instead of loading from CDN. While CDN loading with the fallback works it can be a bit slow as the CDN is probed first before the fallback kicks in. Switching quickly in one place makes this trivial. It also makes it very easy once a new version of jQuery rolls around to move up to the new version and ensure that all pages are using the new version immediately. I'm not trying to make this out as 'the' definite way to load your resources, but rather provide it here as a pointer so you can maybe apply your own logic to determine where scripts come from and how they load. You could even automate this some more by using configuration settings or reading the locations/preferences out of some sort of data/metadata store that can be dynamically updated instead via recompilation. FWIW, I use a very similar approach for loading jQuery UI and my own ww.jquery library - the same concept can be applied to any kind of script you might be loading from different locations. Hopefully some of you find this a useful addition to your toolset. Resources Google CDN for jQuery Full ControlResources Source Code ControlResource Documentation Westwind.Web NuGet This method is part of the Westwind.Web library of the West Wind Web Toolkit or you can grab the Web library from NuGet and add to your Visual Studio project. This package includes a host of Web related utilities and script support features. © Rick Strahl, West Wind Technologies, 2005-2011Posted in ASP.NET  jQuery   Tweet (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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  • Quick guide to Oracle IRM 11g: Configuring SSL

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

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  • SQL SERVER – Securing TRUNCATE Permissions in SQL Server

    - by pinaldave
    Download the Script of this article from here. On December 11, 2010, Vinod Kumar, a Databases & BI technology evangelist from Microsoft Corporation, graced Ahmedabad by spending some time with the Community during the Community Tech Days (CTD) event. As he was running through a few demos, Vinod asked the audience one of the most fundamental and common interview questions – “What is the difference between a DELETE and TRUNCATE?“ Ahmedabad SQL Server User Group Expert Nakul Vachhrajani has come up with excellent solutions of the same. I must congratulate Nakul for this excellent solution and as a encouragement to User Group member, I am publishing the same article over here. Nakul Vachhrajani is a Software Specialist and systems development professional with Patni Computer Systems Limited. He has functional experience spanning legacy code deprecation, system design, documentation, development, implementation, testing, maintenance and support of complex systems, providing business intelligence solutions, database administration, performance tuning, optimization, product management, release engineering, process definition and implementation. He has comprehensive grasp on Database Administration, Development and Implementation with MS SQL Server and C, C++, Visual C++/C#. He has about 6 years of total experience in information technology. Nakul is an member of the Ahmedabad and Gandhinagar SQL Server User Groups, and actively contributes to the community by actively participating in multiple forums and websites like SQLAuthority.com, BeyondRelational.com, SQLServerCentral.com and many others. Please note: The opinions expressed herein are Nakul own personal opinions and do not represent his employer’s view in anyway. All data from everywhere here on Earth go through a series of  four distinct operations, identified by the words: CREATE, READ, UPDATE and DELETE, or simply, CRUD. Putting in Microsoft SQL Server terms, is the process goes like this: INSERT, SELECT, UPDATE and DELETE/TRUNCATE. Quite a few interesting responses were received and evaluated live during the session. To summarize them, the most important similarity that came out was that both DELETE and TRUNCATE participate in transactions. The major differences (not all) that came out of the exercise were: DELETE: DELETE supports a WHERE clause DELETE removes rows from a table, row-by-row Because DELETE moves row-by-row, it acquires a row-level lock Depending upon the recovery model of the database, DELETE is a fully-logged operation. Because DELETE moves row-by-row, it can fire off triggers TRUNCATE: TRUNCATE does not support a WHERE clause TRUNCATE works by directly removing the individual data pages of a table TRUNCATE directly occupies a table-level lock. (Because a lock is acquired, and because TRUNCATE can also participate in a transaction, it has to be a logged operation) TRUNCATE is, therefore, a minimally-logged operation; again, this depends upon the recovery model of the database Triggers are not fired when TRUNCATE is used (because individual row deletions are not logged) Finally, Vinod popped the big homework question that must be critically analyzed: “We know that we can restrict a DELETE operation to a particular user, but how can we restrict the TRUNCATE operation to a particular user?” After returning home and having a nice cup of coffee, I noticed that my gray cells immediately started to work. Below was the result of my research. As what is always said, the devil is in the details. Upon looking at the Permissions section for the TRUNCATE statement in Books On Line, the following jumps right out: “The minimum permission required is ALTER on table_name. TRUNCATE TABLE permissions default to the table owner, members of the sysadmin fixed server role, and the db_owner and db_ddladmin fixed database roles, and are not transferable. However, you can incorporate the TRUNCATE TABLE statement within a module, such as a stored procedure, and grant appropriate permissions to the module using the EXECUTE AS clause.“ Now, what does this mean? Unlike DELETE, one cannot directly assign permissions to a user/set of users allowing or revoking TRUNCATE rights. However, there is a way to circumvent this. It is important to recall that in Microsoft SQL Server, database engine security surrounds the concept of a “securable”, which is any object like a table, stored procedure, trigger, etc. Rights are assigned to a principal on a securable. Refer to the image below (taken from the SQL Server Books On Line). urable”, which is any object like a table, stored procedure, trigger, etc. Rights are assigned to a principal on a securable. Refer to the image below (taken from the SQL Server Books On Line). SETTING UP THE ENVIRONMENT – (01A_Truncate Table Permissions.sql) Script Provided at the end of the article. By the end of this demo, one will be able to do all the CRUD operations, except the TRUNCATE, and the other will only be able to execute the TRUNCATE. All you will need for this test is any edition of SQL Server 2008. (With minor changes, these scripts can be made to work with SQL 2005.) We begin by creating the following: 1.       A test database 2.        Two database roles: associated logins and users 3.       Switch over to the test database and create a test table. Then, add some data into it. I am using row constructors, which is new to SQL 2008. Creating the modules that will be used to enforce permissions 1.       We have already created one of the modules that we will be assigning permissions to. That module is the table: TruncatePermissionsTest 2.       We will now create two stored procedures; one is for the DELETE operation and the other for the TRUNCATE operation. Please note that for all practical purposes, the end result is the same – all data from the table TruncatePermissionsTest is removed Assigning the permissions Now comes the most important part of the demonstration – assigning permissions. A permissions matrix can be worked out as under: To apply the security rights, we use the GRANT and DENY clauses, as under: That’s it! We are now ready for our big test! THE TEST (01B_Truncate Table Test Queries.sql) Script Provided at the end of the article. I will now need two separate SSMS connections, one with the login AllowedTruncate and the other with the login RestrictedTruncate. Running the test is simple; all that’s required is to run through the script – 01B_Truncate Table Test Queries.sql. What I will demonstrate here via screen-shots is the behavior of SQL Server when logged in as the AllowedTruncate user. There are a few other combinations than what are highlighted here. I will leave the reader the right to explore the behavior of the RestrictedTruncate user and these additional scenarios, as a form of self-study. 1.       Testing SELECT permissions 2.       Testing TRUNCATE permissions (Remember, “deny by default”?) 3.       Trying to circumvent security by trying to TRUNCATE the table using the stored procedure Hence, we have now proved that a user can indeed be assigned permissions to specifically assign TRUNCATE permissions. I also hope that the above has sparked curiosity towards putting some security around the probably “destructive” operations of DELETE and TRUNCATE. I would like to wish each and every one of the readers a very happy and secure time with Microsoft SQL Server. (Please find the scripts – 01A_Truncate Table Permissions.sql and 01B_Truncate Table Test Queries.sql that have been used in this demonstration. Please note that these scripts contain purely test-level code only. These scripts must not, at any cost, be used in the reader’s production environments). 01A_Truncate Table Permissions.sql /* ***************************************************************************************************************** Developed By          : Nakul Vachhrajani Functionality         : This demo is focused on how to allow only TRUNCATE permissions to a particular user How to Use            : 1. Run through, step-by-step through the sequence till Step 08 to create a test database 2. Switch over to the "Truncate Table Test Queries.sql" and execute it step-by-step in two different SSMS windows, one where you have logged in as 'RestrictedTruncate', and the other as 'AllowedTruncate' 3. Come back to "Truncate Table Permissions.sql" 4. Execute Step 10 to cleanup! Modifications         : December 13, 2010 - NAV - Updated to add a security matrix and improve code readability when applying security December 12, 2010 - NAV - Created ***************************************************************************************************************** */ -- Step 01: Create a new test database CREATE DATABASE TruncateTestDB GO USE TruncateTestDB GO -- Step 02: Add roles and users to demonstrate the security of the Truncate operation -- 2a. Create the new roles CREATE ROLE AllowedTruncateRole; GO CREATE ROLE RestrictedTruncateRole; GO -- 2b. Create new logins CREATE LOGIN AllowedTruncate WITH PASSWORD = 'truncate@2010', CHECK_POLICY = ON GO CREATE LOGIN RestrictedTruncate WITH PASSWORD = 'truncate@2010', CHECK_POLICY = ON GO -- 2c. Create new Users using the roles and logins created aboave CREATE USER TruncateUser FOR LOGIN AllowedTruncate WITH DEFAULT_SCHEMA = dbo GO CREATE USER NoTruncateUser FOR LOGIN RestrictedTruncate WITH DEFAULT_SCHEMA = dbo GO -- 2d. Add the newly created login to the newly created role sp_addrolemember 'AllowedTruncateRole','TruncateUser' GO sp_addrolemember 'RestrictedTruncateRole','NoTruncateUser' GO -- Step 03: Change over to the test database USE TruncateTestDB GO -- Step 04: Create a test table within the test databse CREATE TABLE TruncatePermissionsTest (Id INT IDENTITY(1,1), Name NVARCHAR(50)) GO -- Step 05: Populate the required data INSERT INTO TruncatePermissionsTest VALUES (N'Delhi'), (N'Mumbai'), (N'Ahmedabad') GO -- Step 06: Encapsulate the DELETE within another module CREATE PROCEDURE proc_DeleteMyTable WITH EXECUTE AS SELF AS DELETE FROM TruncateTestDB..TruncatePermissionsTest GO -- Step 07: Encapsulate the TRUNCATE within another module CREATE PROCEDURE proc_TruncateMyTable WITH EXECUTE AS SELF AS TRUNCATE TABLE TruncateTestDB..TruncatePermissionsTest GO -- Step 08: Apply Security /* *****************************SECURITY MATRIX*************************************** =================================================================================== Object                   | Permissions |                 Login |             | AllowedTruncate   |   RestrictedTruncate |             |User:NoTruncateUser|   User:TruncateUser =================================================================================== TruncatePermissionsTest  | SELECT,     |      GRANT        |      (Default) | INSERT,     |                   | | UPDATE,     |                   | | DELETE      |                   | -------------------------+-------------+-------------------+----------------------- TruncatePermissionsTest  | ALTER       |      DENY         |      (Default) -------------------------+-------------+----*/----------------+----------------------- proc_DeleteMyTable | EXECUTE | GRANT | DENY -------------------------+-------------+-------------------+----------------------- proc_TruncateMyTable | EXECUTE | DENY | GRANT -------------------------+-------------+-------------------+----------------------- *****************************SECURITY MATRIX*************************************** */ /* Table: TruncatePermissionsTest*/ GRANT SELECT, INSERT, UPDATE, DELETE ON TruncateTestDB..TruncatePermissionsTest TO NoTruncateUser GO DENY ALTER ON TruncateTestDB..TruncatePermissionsTest TO NoTruncateUser GO /* Procedure: proc_DeleteMyTable*/ GRANT EXECUTE ON TruncateTestDB..proc_DeleteMyTable TO NoTruncateUser GO DENY EXECUTE ON TruncateTestDB..proc_DeleteMyTable TO TruncateUser GO /* Procedure: proc_TruncateMyTable*/ DENY EXECUTE ON TruncateTestDB..proc_TruncateMyTable TO NoTruncateUser GO GRANT EXECUTE ON TruncateTestDB..proc_TruncateMyTable TO TruncateUser GO -- Step 09: Test --Switch over to the "Truncate Table Test Queries.sql" and execute it step-by-step in two different SSMS windows: --    1. one where you have logged in as 'RestrictedTruncate', and --    2. the other as 'AllowedTruncate' -- Step 10: Cleanup sp_droprolemember 'AllowedTruncateRole','TruncateUser' GO sp_droprolemember 'RestrictedTruncateRole','NoTruncateUser' GO DROP USER TruncateUser GO DROP USER NoTruncateUser GO DROP LOGIN AllowedTruncate GO DROP LOGIN RestrictedTruncate GO DROP ROLE AllowedTruncateRole GO DROP ROLE RestrictedTruncateRole GO USE MASTER GO DROP DATABASE TruncateTestDB GO 01B_Truncate Table Test Queries.sql /* ***************************************************************************************************************** Developed By          : Nakul Vachhrajani Functionality         : This demo is focused on how to allow only TRUNCATE permissions to a particular user How to Use            : 1. Switch over to this from "Truncate Table Permissions.sql", Step #09 2. Execute this step-by-step in two different SSMS windows a. One where you have logged in as 'RestrictedTruncate', and b. The other as 'AllowedTruncate' 3. Return back to "Truncate Table Permissions.sql" 4. Execute Step 10 to cleanup! Modifications         : December 12, 2010 - NAV - Created ***************************************************************************************************************** */ -- Step 09A: Switch to the test database USE TruncateTestDB GO -- Step 09B: Ensure that we have valid data SELECT * FROM TruncatePermissionsTest GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 1 -- The SELECT permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. --Step 09C: Attempt to Truncate Data from the table without using the stored procedure TRUNCATE TABLE TruncatePermissionsTest GO -- (Expected: Following error will occur) --  Msg 1088, Level 16, State 7, Line 2 --  Cannot find the object "TruncatePermissionsTest" because it does not exist or you do not have permissions. -- Step 09D:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'London'), (N'Paris'), (N'Berlin') GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 1 -- The INSERT permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. --Step 09E: Attempt to Truncate Data from the table using the stored procedure EXEC proc_TruncateMyTable GO -- (Expected: Will execute successfully with 'AllowedTruncate' user, will error out as under with 'RestrictedTruncate') -- Msg 229, Level 14, State 5, Procedure proc_TruncateMyTable, Line 1 -- The EXECUTE permission was denied on the object 'proc_TruncateMyTable', database 'TruncateTestDB', schema 'dbo'. -- Step 09F:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'Madrid'), (N'Rome'), (N'Athens') GO --Step 09G: Attempt to Delete Data from the table without using the stored procedure DELETE FROM TruncatePermissionsTest GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 2 -- The DELETE permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. -- Step 09H:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'Spain'), (N'Italy'), (N'Greece') GO --Step 09I: Attempt to Delete Data from the table using the stored procedure EXEC proc_DeleteMyTable GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Procedure proc_DeleteMyTable, Line 1 -- The EXECUTE permission was denied on the object 'proc_DeleteMyTable', database 'TruncateTestDB', schema 'dbo'. --Step 09J: Close this SSMS window and return back to "Truncate Table Permissions.sql" Thank you Nakul to take up the challenge and prove that Ahmedabad and Gandhinagar SQL Server User Group has talent to solve difficult problems. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Best Practices, Pinal Dave, Readers Contribution, Readers Question, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Security, SQL Server, SQL Tips and Tricks, T SQL, Technology

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

    - by dwahlin
    Technology is fun isn’t it? As soon as you think you’ve figured out where things are heading a new technology comes onto the scene, changes things up, and offers new opportunities. One of the new technologies I’ve been spending quite a bit of time with lately is Windows 8 store applications. I posted my thoughts about Windows 8 during the BUILD conference in 2011 and still feel excited about the opportunity there. Time will tell how well it ends up being accepted by consumers but I’m hopeful that it’ll take off. I currently have two Windows 8 store application concepts I’m working on with one being built in XAML/C# and another in HTML/JavaScript. I really like that Microsoft supports both options since it caters to a variety of developers and makes it easy to get started regardless if you’re a desktop developer or Web developer. Here’s a quick look at how the technologies are organized in Windows 8: In this post I’ll focus on the basics of Windows 8 store XAML/C# apps by looking at features, files, and code provided by Visual Studio projects. To get started building these types of apps you’ll definitely need to have some knowledge of XAML and C#. Let’s get started by looking at the Windows 8 store project types available in Visual Studio 2012.   Windows 8 Store XAML/C# Project Types When you open Visual Studio 2012 you’ll see a new entry under C# named Windows Store. It includes 6 different project types as shown next.   The Blank App project provides initial starter code and a single page whereas the Grid App and Split App templates provide quite a bit more code as well as multiple pages for your application. The other projects available can be be used to create a class library project that runs in Windows 8 store apps, a WinRT component such as a custom control, and a unit test library project respectively. If you’re building an application that displays data in groups using the “tile” concept then the Grid App or Split App project templates are a good place to start. An example of the initial screens generated by each project is shown next: Grid App Split View App   When a user clicks a tile in a Grid App they can view details about the tile data. With a Split View app groups/categories are shown and when the user clicks on a group they can see a list of all the different items and then drill-down into them:   For the remainder of this post I’ll focus on functionality provided by the Blank App project since it provides a simple way to get started learning the fundamentals of building Windows 8 store apps.   Blank App Project Walkthrough The Blank App project is a great place to start since it’s simple and lets you focus on the basics. In this post I’ll focus on what it provides you out of the box and cover additional details in future posts. Once you have the basics down you can move to the other project types if you need the functionality they provide. The Blank App project template does exactly what it says – you get an empty project with a few starter files added to help get you going. This is a good option if you’ll be building an app that doesn’t fit into the grid layout view that you see a lot of Windows 8 store apps following (such as on the Windows 8 start screen). I ended up starting with the Blank App project template for the app I’m currently working on since I’m not displaying data/image tiles (something the Grid App project does well) or drilling down into lists of data (functionality that the Split App project provides). The Blank App project provides images for the tiles and splash screen (you’ll definitely want to change these), a StandardStyles.xaml resource dictionary that includes a lot of helpful styles such as buttons for the AppBar (a special type of menu in Windows 8 store apps), an App.xaml file, and the app’s main page which is named MainPage.xaml. It also adds a Package.appxmanifest that is used to define functionality that your app requires, app information used in the store, plus more. The App.xaml, App.xaml.cs and StandardStyles.xaml Files The App.xaml file handles loading a resource dictionary named StandardStyles.xaml which has several key styles used throughout the application: <Application x:Class="BlankApp.App" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:local="using:BlankApp"> <Application.Resources> <ResourceDictionary> <ResourceDictionary.MergedDictionaries> <!-- Styles that define common aspects of the platform look and feel Required by Visual Studio project and item templates --> <ResourceDictionary Source="Common/StandardStyles.xaml"/> </ResourceDictionary.MergedDictionaries> </ResourceDictionary> </Application.Resources> </Application>   StandardStyles.xaml has style definitions for different text styles and AppBar buttons. If you scroll down toward the middle of the file you’ll see that many AppBar button styles are included such as one for an edit icon. Button styles like this can be used to quickly and easily add icons/buttons into your application without having to be an expert in design. <Style x:Key="EditAppBarButtonStyle" TargetType="ButtonBase" BasedOn="{StaticResource AppBarButtonStyle}"> <Setter Property="AutomationProperties.AutomationId" Value="EditAppBarButton"/> <Setter Property="AutomationProperties.Name" Value="Edit"/> <Setter Property="Content" Value="&#xE104;"/> </Style> Switching over to App.xaml.cs, it includes some code to help get you started. An OnLaunched() method is added to handle creating a Frame that child pages such as MainPage.xaml can be loaded into. The Frame has the same overall purpose as the one found in WPF and Silverlight applications - it’s used to navigate between pages in an application. /// <summary> /// Invoked when the application is launched normally by the end user. Other entry points /// will be used when the application is launched to open a specific file, to display /// search results, and so forth. /// </summary> /// <param name="args">Details about the launch request and process.</param> protected override void OnLaunched(LaunchActivatedEventArgs args) { Frame rootFrame = Window.Current.Content as Frame; // Do not repeat app initialization when the Window already has content, // just ensure that the window is active if (rootFrame == null) { // Create a Frame to act as the navigation context and navigate to the first page rootFrame = new Frame(); if (args.PreviousExecutionState == ApplicationExecutionState.Terminated) { //TODO: Load state from previously suspended application } // Place the frame in the current Window Window.Current.Content = rootFrame; } if (rootFrame.Content == null) { // When the navigation stack isn't restored navigate to the first page, // configuring the new page by passing required information as a navigation // parameter if (!rootFrame.Navigate(typeof(MainPage), args.Arguments)) { throw new Exception("Failed to create initial page"); } } // Ensure the current window is active Window.Current.Activate(); }   Notice that in addition to creating a Frame the code also checks to see if the app was previously terminated so that you can load any state/data that the user may need when the app is launched again. If you’re new to the lifecycle of Windows 8 store apps the following image shows how an app can be running, suspended, and terminated.   If the user switches from an app they’re running the app will be suspended in memory. The app may stay suspended or may be terminated depending on how much memory the OS thinks it needs so it’s important to save state in case the application is ultimately terminated and has to be started fresh. Although I won’t cover saving application state here, additional information can be found at http://msdn.microsoft.com/en-us/library/windows/apps/xaml/hh465099.aspx. Another method in App.xaml.cs named OnSuspending() is also included in App.xaml.cs that can be used to store state as the user switches to another application:   /// <summary> /// Invoked when application execution is being suspended. Application state is saved /// without knowing whether the application will be terminated or resumed with the contents /// of memory still intact. /// </summary> /// <param name="sender">The source of the suspend request.</param> /// <param name="e">Details about the suspend request.</param> private void OnSuspending(object sender, SuspendingEventArgs e) { var deferral = e.SuspendingOperation.GetDeferral(); //TODO: Save application state and stop any background activity deferral.Complete(); } The MainPage.xaml and MainPage.xaml.cs Files The Blank App project adds a file named MainPage.xaml that acts as the initial screen for the application. It doesn’t include anything aside from an empty <Grid> XAML element in it. The code-behind class named MainPage.xaml.cs includes a constructor as well as a method named OnNavigatedTo() that is called once the page is displayed in the frame.   /// <summary> /// An empty page that can be used on its own or navigated to within a Frame. /// </summary> public sealed partial class MainPage : Page { public MainPage() { this.InitializeComponent(); } /// <summary> /// Invoked when this page is about to be displayed in a Frame. /// </summary> /// <param name="e">Event data that describes how this page was reached. The Parameter /// property is typically used to configure the page.</param> protected override void OnNavigatedTo(NavigationEventArgs e) { } }   If you’re experienced with XAML you can switch to Design mode and start dragging and dropping XAML controls from the ToolBox in Visual Studio. If you prefer to type XAML you can do that as well in the XAML editor or while in split mode. Many of the controls available in WPF and Silverlight are included such as Canvas, Grid, StackPanel, and Border for layout. Standard input controls are also included such as TextBox, CheckBox, PasswordBox, RadioButton, ComboBox, ListBox, and more. MediaElement is available for rendering video or playing audio files. Some of the “common” XAML controls included out of the box are shown next:   Although XAML/C# Windows 8 store apps don’t include all of the functionality available in Silverlight 5, the core functionality required to build store apps is there with additional functionality available in open source projects such as Callisto (started by Microsoft’s Tim Heuer), Q42.WinRT, and others. Standard XAML data binding can be used to bind C# objects to controls, converters can be used to manipulate data during the data binding process, and custom styles and templates can be applied to controls to modify them. Although Visual Studio 2012 doesn’t support visually creating styles or templates, Expression Blend 5 handles that very well. To get started building the initial screen of a Windows 8 app you can start adding controls as mentioned earlier. Simply place them inside of the <Grid> element that’s included. You can arrange controls in a stacked manner using the StackPanel control, add a border around controls using the Border control, arrange controls in columns and rows using the Grid control, or absolutely position controls using the Canvas control. One of the controls that may be new to you is the AppBar. It can be used to add menu/toolbar functionality into a store app and keep the app clean and focused. You can place an AppBar at the top or bottom of the screen. A user on a touch device can swipe up to display the bottom AppBar or right-click when using a mouse. An example of defining an AppBar that contains an Edit button is shown next. The EditAppBarButtonStyle is available in the StandardStyles.xaml file mentioned earlier. <Page.BottomAppBar> <AppBar x:Name="ApplicationAppBar" Padding="10,0,10,0" AutomationProperties.Name="Bottom App Bar"> <Grid> <StackPanel x:Name="RightPanel" Orientation="Horizontal" Grid.Column="1" HorizontalAlignment="Right"> <Button x:Name="Edit" Style="{StaticResource EditAppBarButtonStyle}" Tag="Edit" /> </StackPanel> </Grid> </AppBar> </Page.BottomAppBar> Like standard XAML controls, the <Button> control in the AppBar can be wired to an event handler method in the MainPage.Xaml.cs file or even bound to a ViewModel object using “commanding” if your app follows the Model-View-ViewModel (MVVM) pattern (check out the MVVM Light package available through NuGet if you’re using MVVM with Windows 8 store apps). The AppBar can be used to navigate to different screens, show and hide controls, display dialogs, show settings screens, and more.   The Package.appxmanifest File The Package.appxmanifest file contains configuration details about your Windows 8 store app. By double-clicking it in Visual Studio you can define the splash screen image, small and wide logo images used for tiles on the start screen, orientation information, and more. You can also define what capabilities the app has such as if it uses the Internet, supports geolocation functionality, requires a microphone or webcam, etc. App declarations such as background processes, file picker functionality, and sharing can also be defined Finally, information about how the app is packaged for deployment to the store can also be defined. Summary If you already have some experience working with XAML technologies you’ll find that getting started building Windows 8 applications is pretty straightforward. Many of the controls available in Silverlight and WPF are available making it easy to get started without having to relearn a lot of new technologies. In the next post in this series I’ll discuss additional features that can be used in your Windows 8 store apps.

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  • Mobile Friendly Websites with CSS Media Queries

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

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