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  • Should I be using a JavaScript SPA designed when security is important

    - by ryanzec
    I asked something kind of similar on stackoverflow with a particular piece of code however I want to try to ask this in a broader sense. So I have this web application that I have started to write in backbone using a Single Page Architecture (SPA) however I am starting to second guess myself because of security. Now we are not storing and sending credit card information or anything like that through this web application but we are storing sensitive information that people are uploading to us and will have the ability to re-download too. The obviously security concern that I have with JavaScript is that you can't trust anything that comes from JavaScript however in a Backbone SPA application, everything is being sent through JavaScript. There are two security features that I will have to build in JavaScript; permissions and authentication. The authentication piece is just me override the Backbone.Router.prototype.navigate method to check the fragment it is trying to load and if the JavaScript application.session.loggedIn is not set to true (and they are not viewing a none authenticated page), they are redirected to the login page automatically. The user could easily modify application.session.loggedIn to equal true (or modify Backbone.Router.prototype.navigate method) but then they would also have to not so easily dynamically embedded a link into the page (or modify a current one) that has the proper classes, data-* attributes, and href values to then load a page that should only be loaded when they user has logged in (and has the permissions). So I have an acl object that deals with the permissions stuff. All someone would have to do to view pages or parts of pages they should not be able to is to call acl.addPermission(resource, permission) with the proper permissions or modify the acl.hasPermission() to always return true and then navigate away and then back to the page. Now certain things is EMCAScript 5 like Object.seal() or Object.freeze() would help with some of this however we have to support IE 8 which does not support those pieces of functionality. Now the REST API also performs security checks on every request so technically even if they are able to see parts of the interface that they should not be able to, they still should not be able to actually affect any data. The main benefits for me in developing a JavaScript SPA application is that the application is a lot more responsive since it is only transferring the minimum amount of JSON data for the requested action and performing the minimum amount of work too. There are also other things that I think are beneficial like you are going to have to develop an API for the data (which is good if you want expand your application to different platforms/technologies) or their is more of a separation between front-end and back-end however if security is a concern, it is really wise to go down the road of a JavaScript SPA application for the front-end?

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  • Using MVP, how to create a view from another view, linked with the same model object

    - by Dinaiz
    Background We use the Model-View-Presenter design pattern along with the abstract factory pattern and the "signal/slot" pattern in our application, to fullfill 2 main requirements Enhance testability (very lightweight GUI, every action can be simulated in unit tests) Make the "view" totally independant from the rest, so we can change the actual view implementation, without changing anything else In order to do so our code is divided in 4 layers : Core : which holds the model Presenter : which manages interactions between the view interfaces (see bellow) and the core View Interfaces : they define the signals and slots for a View, but not the implementation Views : the actual implementation of the views When the presenter creates or deals with views, it uses an abstract factory and only knows about the view interfaces. It does the signal/slot binding between views interfaces. It doesn't care about the actual implementation. In the "views" layer, we have a concrete factory which deals with implementations. The signal/slot mechanism is implemented using a custom framework built upon boost::function. Really, what we have is something like that : http://martinfowler.com/eaaDev/PassiveScreen.html Everything works fine. The problem However, there's a problem I don't know how to solve. Let's take for example a very simple drag and drop example. I have two ContainersViews (ContainerView1, ContainerView2). ContainerView1 has an ItemView1. I drag the ItemView1 from ContainerView1 to ContainerView2. ContainerView2 must create an ItemView2, of a different type, but which "points" to the same model object as ItemView1. So the ContainerView2 gets a callback called for the drop action with ItemView1 as a parameter. It calls ContainerPresenterB passing it ItemViewB In this case we are only dealing with views. In MVP-PV, views aren't supposed to know anything about the presenter nor the model, right ? How can I create the ItemView2 from the ItemView1, not knowing which model object is ItemView1 representing ? I thought about adding an "itemId" to every view, this id being the id of the core object the view represents. So in pseudo code, ContainerPresenter2 would do something like itemView2=abstractWidgetFactory.createItemView2(); this.add(itemView2,itemView1.getCoreObjectId()) I don't get too much into details. That just work. The problem I have here is that those itemIds are just like pointers. And pointers can be dangling. Imagine that by mistake, I delete itemView1, and this deletes coreObject1. The itemView2 will have a coreObjectId which represents an invalid coreObject. Isn't there a more elegant and "bulletproof" solution ? Even though I never did ObjectiveC or macOSX programming, I couldn't help but notice that our framework is very similar to Cocoa framework. How do they deal with this kind of problem ? Couldn't find more in-depth information about that on google. If someone could shed some light on this. I hope this question isn't too confusing ...

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  • Using SQL Source Control and Vault Professional Part 4

    - by Ajarn Mark Caldwell
    Two weeks ago I upgraded our installation of Fortress to the latest version, which is now named Vault Professional.  This is the version of Vault (i.e. Vault Standard 5.1 / Vault Professional 5.1) that will be officially supported with Red-Gate SQL Source Control 2.1.  While the folks at Red-Gate did a fantastic job of working with me to get SQL Source Control to work with the older Fortress version, we weren’t going to just sit on that.  There are a couple of things that Vault Professional cleaned up for us, such as improved integration with Visual Studio 2010, so it was a win all around. Shortly after that upgrade, I received notice from Red-Gate that they had a new Early Access version of SQL Source Control available that included the ability to source control static data.  The idea here is that you probably have a few fairly static lookup tables in your system, and those data values are similar in concept to source code, and should be versioned in your source control management system also.  I agree with this, but please be wise…somebody out there is bound to try to use this feature as their disaster recovery for their entire database, and that is NOT the purpose.  First off, you should never have your PROD (or LIVE, whatever you call it) system attached to source control.  Source Control is for development, not for PROD systems.  Second, use the features that are intended for this purpose, such as BACKUP and RESTORE. Laying that tangent aside, it is great that now you can include these critical values in your repository and make them part of a deployment process.  As you would guess, SQL Source Control uses SQL Data Compare to create the data change scripts just like it uses SQL Compare to create the schema change scripts.  Once again, they did a very good job with the integration to their other products.  At this point we are really starting to see some good payback on our investment in the full SQL Developer Bundle.  Those products were worth the investment back when we only used them sporadically for troubleshooting and DBA analysis, but now with SQL Source Control, they are becoming everyday-use products for the development team. I like this software (SQL Source Control) so much that I am about to break my own rules and distribute it to my team to use even though it is still in beta.  This is the first time that I have approved the use of any beta software in a production scenario (actively building our next versions of internal software) but I predict that the usability and productivity gain of using SQL Source Control over manual scripting is worth the risk.  Of course, I have also put this beta software through its paces pretty well to be comfortable with it, and Red-Gate has proven their responsiveness to issues that came up in my early beta testing, and so I am willing to bet on their continued support.  Likewise, SourceGear, the maker of Vault Professional, has proven itself to me as well, and so the combination of SQL Source Control with Vault Professional is the new standard for my development team.

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  • Using multiple column layout with HTML 5 and CSS 3

    - by nikolaosk
    This is going to be the fourth post in a series of posts regarding HTML 5. You can find the other posts here , here and here.In this post I will provide a hands-on example with HTML 5 and CSS 3 on how to create a page with multiple columns and proper layout.I will show you how to use CSS 3 to create columns much easier than relying on DIV elements and the float CSS rule.I will also show you how to use browser-specific prefix rules (-ms for Internet Explorer and -moz for Firefox ) for browsers that do not fully support CSS 3.In order to be absolutely clear this is not (and could not be) a detailed tutorial on HTML 5. There are other great resources for that.Navigate to the excellent interactive tutorials of W3School.Another excellent resource is HTML 5 Doctor.Two very nice sites that show you what features and specifications are implemented by various browsers and their versions are http://caniuse.com/ and http://html5test.com/. At this times Chrome seems to support most of HTML 5 specifications.Another excellent way to find out if the browser supports HTML 5 and CSS 3 features is to use the Javascript lightweight library Modernizr.In this hands-on example I will be using Expression Web 4.0.This application is not a free application. You can use any HTML editor you like.You can use Visual Studio 2012 Express edition. You can download it here.I will create a simple page with information about HTML 5, CSS 3 and JQuery. This is the full HTML 5 code. <!DOCTYPE html><html lang="en">  <head>    <title>HTML 5, CSS3 and JQuery</title>    <meta http-equiv="Content-Type" content="text/html;charset=utf-8" >    <link rel="stylesheet" type="text/css" href="style.css">       </head>  <body>    <div id="header">      <h1>Learn cutting edge technologies</h1>      <p>HTML 5, JQuery, CSS3</p>    </div>    <div id="main">      <div id="mainnews">        <div>          <h2>HTML 5</h2>        </div>        <div>          <p>            HTML5 is the latest version of HTML and XHTML. The HTML standard defines a single language that can be written in HTML and XML. It attempts to solve issues found in previous iterations of HTML and addresses the needs of Web Applications, an area previously not adequately covered by HTML.          </p>          <div class="quote">            <h4>Do More with Less</h4>            <p>             jQuery is a fast and concise JavaScript Library that simplifies HTML document traversing, event handling, animating, and Ajax interactions for rapid web development.             </p>            </div>          <p>            The HTML5 test(html5test.com) score is an indication of how well your browser supports the upcoming HTML5 standard and related specifications. Even though the specification isn't finalized yet, all major browser manufacturers are making sure their browser is ready for the future. Find out which parts of HTML5 are already supported by your browser today and compare the results with other browsers.                      The HTML5 test does not try to test all of the new features offered by HTML5, nor does it try to test the functionality of each feature it does detect. Despite these shortcomings we hope that by quantifying the level of support users and web developers will get an idea of how hard the browser manufacturers work on improving their browsers and the web as a development platform.</p>        </div>      </div>              <div id="CSS">        <div>          <h2>CSS 3 Intro</h2>        </div>        <div>          <p>          Cascading Style Sheets (CSS) is a style sheet language used for describing the presentation semantics (the look and formatting) of a document written in a markup language. Its most common application is to style web pages written in HTML and XHTML, but the language can also be applied to any kind of XML document, including plain XML, SVG and XUL.          </p>        </div>      </div>            <div id="CSSmore">        <div>          <h2>CSS 3 Purpose</h2>        </div>        <div>          <p>            CSS is designed primarily to enable the separation of document content (written in HTML or a similar markup language) from document presentation, including elements such as the layout, colors, and fonts.[1] This separation can improve content accessibility, provide more flexibility and control in the specification of presentation characteristics, enable multiple pages to share formatting, and reduce complexity and repetition in the structural content (such as by allowing for tableless web design).          </p>        </div>      </div>                </div>    <div id="footer">        <p>Feel free to google more about the subject</p>      </div>     </body>  </html>  The markup is very easy to follow. I have used some HTML 5 tags and the relevant HTML 5 doctype.The CSS code (style.css) follows  body{        line-height: 30px;        width: 1024px;        background-color:#eee;      }            p{        font-size:17px;    font-family:"Comic Sans MS"      }      p,h2,h3,h4{        margin: 0 0 20px 0;      }            #main, #header, #footer{        width: 100%;        margin: 0px auto;        display:block;      }            #header{        text-align: center;         border-bottom: 1px solid #000;         margin-bottom: 30px;      }            #footer{        text-align: center;         border-top: 1px solid #000;         margin-bottom: 30px;      }            .quote{        width: 200px;       margin-left: 10px;       padding: 5px;       float: right;       border: 2px solid #000;       background-color:#F9ACAE;      }            .quote :last-child{        margin-bottom: 0;      }            #main{        column-count:2;        column-gap:20px;        column-rule: 1px solid #000;        -moz-column-count: 2;        -webkit-column-count: 2;        -moz-column-gap: 20px;        -webkit-column-gap: 20px;        -moz-column-rule: 1px solid #000;        -webkit-column-rule: 1px solid #000;      }       All the rules in the css code are pretty simple. The layout is achieved with that CSS rule #main{        column-count:2;        column-gap:20px;        column-rule: 1px solid #000;        -moz-column-count: 2;        -webkit-column-count: 2;        -moz-column-gap: 20px;        -webkit-column-gap: 20px;        -moz-column-rule: 1px solid #000;        -webkit-column-rule: 1px solid #000; Do note the column-count,column-gap and column-rule properties. These properties make the two column layout possible.Please have a look at the picture below to see why I used prefixes for Chrome (webkit) and Firefox(moz).It clearly indicates that the CSS 3 column layout are not supported from Firefox and Chrome.   Finally I test my simple HTML 5 page using the latest versions of Firefox,Internet Explorer and Chrome. In my machine I have installed Firefox 15.0.1.Have a look at the picture below to see how the page looks  I have installed Google Chrome 21.0 in my machine.Have a look at the picture below to see how the page looks Have a look at the picture below to see how my page looks in IE 10.  My page looks the same in all browsers. Hope it helps!!!

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  • Using Deployment Manager

    - by Jess Nickson
    One of the teams at Red Gate has been working very hard on a new product: Deployment Manager. Deployment Manager is a free tool that lets you deploy updates to .NET apps, services and databases through a central dashboard. Deployment Manager has been out for a while, but I must admit that even though I work in the same building, until now I hadn’t even looked at it. My job at Red Gate is to develop and maintain some of our community sites, which involves carrying out regular deployments. One of the projects I have to deploy on a fairly regular basis requires me to send my changes to our build server, TeamCity. The output is a Zip file of the build. I then have to go and find this file, copy it across to the staging machine, extract it, and copy some of the sub-folders to other places. In order to keep track of what builds are running, I need to rename the folders accordingly. However, even after all that, I still need to go and update the site and its applications in IIS to point at these new builds. Oh, and then, I have to repeat the process when I deploy on production. Did I mention the multiple configuration files that then need updating as well? Manually? The whole process can take well over half an hour. I’m ready to try out a new process. Deployment Manager is designed to massively simplify the deployment processes from what could be lots of manual copying of files, managing of configuration files, and database upgrades down to a few clicks. It’s a big promise, but I decided to try out this new tool on one of the smaller ASP.NET sites at Red Gate, Format SQL (the result of a Red Gate Down Tools week). I wanted to add some new functionality, but given it was a new site with no set way of doing things, I was reluctant to have to manually copy files around servers. I decided to use this opportunity as a chance to set the site up on Deployment Manager and check out its functionality. What follows is a guide on how to get set up with Deployment Manager, a brief overview of its features, and what I thought of the experience. To follow along with the instructions that follow, you’ll first need to download Deployment Manager from Red Gate. It has a free ‘Starter Edition’ which allows you to create up to 5 projects and agents (machines you deploy to), so it’s really easy to get up and running with a fully-featured version. The Initial Set Up After installing the product and setting it up using the administration tool it provides, I launched Deployment Manager by going to the URL and port I had set it to run on. This loads up the main dashboard. The dashboard does a good job of guiding me through the process of getting started, beginning with a prompt to create some environments. 1. Setting up Environments The dashboard informed me that I needed to add new ‘Environments’, which are essentially ways of grouping the machines you want to deploy to. The environments that get added will show up on the main dashboard. I set up two such environments for this project: ‘staging’ and ‘live’.   2. Add Target Machines Once I had created the environments, I was ready to add ‘target machine’s to them, which are the actual machines that the deployment will occur on.   To enable me to deploy to a new machine, I needed to download and install an Agent on it. The ‘Add target machine’ form on the ‘Environments’ page helpfully provides a link for downloading an Agent.   Once the agent has been installed, it is just a case of copying the server key to the agent, and the agent key to the server, to link them up.   3. Run Health Check If, after adding your new target machine, the ‘Status’ flags an error, it is possible that the Agent and Server keys have not been entered correctly on both Deployment Manager and the Agent service.     You can ‘Check Health’, which will give you more information on any issues. It is probably worth running this regardless of what status the ‘Environments’ dashboard is claiming, just to be on the safe side.     4. Add Projects Going back to the main Dashboard tab at this point, I found that it was telling me that I needed to set up a new project.   I clicked the ‘project’ link to get started, gave my new project a name and clicked ‘Create’. I was then redirected to the ‘Steps’ page for the project under the Projects tab.   5. Package Steps The ‘Steps’ page was fairly empty when it first loaded.   Adding a ‘step’ allowed me to specify what packages I wanted to grab for the deployment. This part requires a NuGet package feed to be set up, which is where Deployment Manager will look for the packages. At Red Gate, we already have one set up, so I just needed to tell Deployment Manager about it. Don’t worry; there is a nice guide included on how to go about doing all of this on the ‘Package Feeds’ page in ‘Settings’, if you need any help with setting these bits up.    At Red Gate we use a build server, TeamCity, which is capable of publishing built projects to the NuGet feed we use. This makes the workflow for Format SQL relatively simple: when I commit a change to the project, the build server is configured to grab those changes, build the project, and spit out a new NuGet package to the Red Gate NuGet package feed. My ‘package step’, therefore, is set up to look for this package on our feed. The final part of package step was simply specifying which machines from what environments I wanted to be able to deploy the project to.     Format SQL Now the main Dashboard showed my new project and environment in a rather empty looking grid. Clicking on my project presented me with a nice little message telling me that I am now ready to create my first release!   Create a release Next I clicked on the ‘Create release’ button in the Projects tab. If your feeds and package step(s) were set up correctly, then Deployment Manager will automatically grab the latest version of the NuGet package that you want to deploy. As you can see here, it was able to pick up the latest build for Format SQL and all I needed to do was enter a version number and description of the release.   As you can see underneath ‘Version number’, it keeps track of what version the previous release was given. Clicking ‘Create’ created the release and redirected me to a summary of it where I could check the details before deploying.   I clicked ‘Deploy this release’ and chose the environment I wanted to deploy to and…that’s it. Deployment Manager went off and deployed it for me.   Once I clicked ‘Deploy release’, Deployment Manager started to automatically update and provide continuing feedback about the process. If any errors do arise, then I can expand the results to see where it went wrong. That’s it, I’m done! Keep in mind, if you hit errors with the deployment itself then it is possible to view the log output to try and determine where these occurred. You can keep expanding the logs to narrow down the problem. The screenshot below is not from my Format SQL deployment, but I thought I’d post one to demonstrate the logging output available. Features One of the best bits of Deployment Manager for me is the ability to very, very easily deploy the same release to multiple machines. Deploying this same release to production was just a case of selecting the deployment and choosing the ‘live’ environment as the place to deploy to. Following on from this is the fact that, as Deployment Manager keeps track of all of your releases, it is extremely easy to roll back to a previous release if anything goes pear-shaped! You can view all your previous releases and select one to re-deploy. I needed this feature more than once when differences in my production and staging machines lead to some odd behavior.     Another option is to use the TeamCity integration available. This enables you to set Deployment Manager up so that it will automatically create releases and deploy these to an environment directly from TeamCity, meaning that you can always see the latest version up and running without having to do anything. Machine Specific Deployments ‘What about custom configuration files?’ I hear you shout. Certainly, it was one of my concerns. Our setup on the staging machine is not in line with that on production. What this means is that, should we deploy the same configuration to both, one of them is going to break. Thankfully, it turns out that Deployment Manager can deal with this. Given I had environments ‘staging’ and ‘live’, and that staging used the project’s web.config file, while production (‘live’) required the config file to undergo some transformations, I simply added a web.live.config file in the project, so that it would be included as part of the NuGet package. In this file, I wrote the XML document transformations I needed and Deployment Manager took care of the rest. Another option is to set up ‘variables’ for your project, which allow you to specify key-value pairs for your configuration file, and which environment to apply them to. You’ll find Variables as a full left-hand submenu within the ‘Projects’ tab. These features will definitely be of interest if you have a large number of environments! There are still many other features that I didn’t get a chance to play around with like running PowerShell scripts for more personalised deployments. Maybe next time! Also, let’s not forget that my use case in this article is a very simple one – deploying a single package. I don’t believe that all projects will be equally as simple, but I already appreciate how much easier Deployment Manager could make my life. I look forward to the possibility of moving our other sites over to Deployment Manager in the near future.   Conclusion In this article I have described the steps involved in setting up and configuring an instance of Deployment Manager, creating a new automated deployment process, and using this to actually carry out a deployment. I’ve tried to mention some of the features I found particularly useful, such as error logging, easy release management allowing you to deploy the same release multiple times, and configuration file transformations. If I had to point out one issue, then it would be that the releases are immutable, which from a development point of view makes sense. However, this causes confusion where I have to create a new release to deploy to a newly set up environment – I cannot simply deploy an old release onto a new environment, the whole release needs to be recreated. I really liked how easy it was to get going with the product. Setting up Format SQL and making a first deployment took very little time. Especially when you compare it to how long it takes me to manually deploy the other site, as I described earlier. I liked how it let me know what I needed to do next, with little messages flagging up that I needed to ‘create environments’ or ‘add some deployment steps’ before I could continue. I found the dashboard incredibly convenient. As the number of projects and environments increase, it might become awkward to try and search them and find out what state they are in. Instead, the dashboard handily keeps track of the latest deployments of each project and lets you know what version is running on each of the environments, and when that deployment occurred. Finally, do you remember my complaint about having to rename folders so that I could keep track of what build they came from? This is yet another thing that Deployment Manager takes care of for you. Each release is put into its own directory, which takes the name of whatever version number that release has, though these can be customised if necessary. If you’d like to take a look at Deployment Manager for yourself, then you can download it here.

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  • Tips On Using The Service Contracts Import Program

    - by LuciaC
    Prior to release 12.1 there was no supported way to import contracts into the EBS Service Contracts application - there were no public APIs nor contract load programs provided.  From release 12.1 onwards the 'Service Contracts Import Program' is provided to load service contracts into the application. The Service Contracts Import functionality is explained in How to Use the Service Contracts Import Program - Scope and Limitations (Doc ID 1057242.1).  This note includes an attached document which explains the program architecture, shows the Entity Relationship Diagram and details the interface table definitions. The Import program takes data from the interface tables listed below and populates the contracts schema tables:  OKS_USAGE_COUNTERS_INTERFACE OKS_SALES_CREDITS_INTERFACEOKS_NOTES_INTERFACEOKS_LINES_INTERFACEOKS_HEADERS_INTERFACEOKS_COVERED_LEVELS_INTERFACEThese interface tables must be loaded via a custom load program.The Service Contracts Import concurrent request is then submitted to create contracts from this legacy data. The parameters to run the Import program are:  Parameter Description  Mode Validate only, Import  Batch Number Batch_Id (unique id populated into the OKS_HEADERS_INTERFACE table)  Number of Workers Number of workers required (these are spawned as separate sub-requests)  Commit size Represents number of successfully processed contracts commited to database The program spawns sub-requests for the import worker(s) and the 'Service Contracts Import Report'.  The data is validated prior to import and into the Contracts tables and will report errors in the Service Contracts Import Report program output file (Import Execution Report).  Troubleshooting tips are provided in R12.1 - Common Service Contract Import Errors (Doc ID 762545.1); this document lists some, but not all, import errors.  The document will be updated over time.  Additional help is given in Debugging Tip for Service Contracts Import Errors (Doc ID 971426.1).After you successfully import contracts, you can purge the records from the interface tables by running the Service Contracts Import Purge concurrent program. Note that there is no supported way to mass delete data from the Contracts schema tables once they are populated, so data loaded by the Import program must be fully tested and verified before the program is run to load data into a Production system.A Service Contracts Import Test program has been provided which will take an existing contract in the application and load the interface tables using the data from that contract.  This can be used as an example for guidance on how to load the interface tables.  The Test program functionality is explained in How to Use the Service Contracts Test Import Program Provided in Release 12.1 (Doc ID 761209.1).  Note that the Test program has some limitations which do not apply to the full Import program and is not a supported program, it is simply a testing tool.  

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  • Cluster Node Recovery Using Second Node in Solaris Cluster

    - by Onur Bingul
    Assumptions:Node 0a is the cluster node that has crashed and could not boot anymore.Node 0b is the node in cluster and in production with services active.Both nodes have their boot disk mirrored via SDS/SVM.We have many options to clone the boot disk from node 0b:- make a copy via network using the ufsdump command and pipe to ufsrestore - make a copy inserting the disk locally on node 0b and creating the third mirror with SDS- make a copy inserting the disk locally on node 0b using dd commandIn this procedure we are going to use dd command (from my experience this is the best option).Bare in mind that in the examples provided we work on Sun Fire V240 systems which have SCSI internal disks. In the case of Fibre Channel (FC) internal disks you must pay attention to the unique identifier, or World Wide Name (WWN), associated with each FC disk (in this case take a look at infodoc #40133 in order to recreate the device tree correctly).Procedure:On node 0b the boot disk is c1t0d0 (c1t1d0 mirror) and this is the VTOC:* Partition  Tag  Flags    Sector     Count    Sector  Mount Directory      0      2    00          0   2106432   2106431      1      3    01    2106432  74630784  76737215      2      5    00          0 143349312 143349311      4      7    00   76737216  50340672 127077887      5      4    00  127077888  14683968 141761855      6      0    00  141761856   1058304 142820159      7      0    00  142820160    529152 143349311We will insert the new disk on node 0b and it will be seen as c1t2d0.1) On node 0b we make a copy via dd from disk c1t0d0s2 to disk c1t2d0s2# dd if=/dev/rdsk/c1t0d0s2 of=/dev/rdsk/c1t2d0s2 bs=8192kA copy of a 72GB disk will take approximately about 45 minutes.Note: as an alternative to make identical copy of root over network follow Document ID: 47498Title: Sun[TM] Cluster 3.0: How to Rebuild a node with Veritas Volume Manager2) Perform an fsck on disk c1t2d0 data slices:   1.  fsck -o f /dev/rdsk/c1t2d0s0 (root)   2.  fsck -o f /dev/rdsk/c1t2d0s4 (/var)   3.  fsck -o f /dev/rdsk/c1t2d0s5 (/usr)   4.  fsck -o f /dev/rdsk/c1t2d0s6 (/globaldevices)3) Mount the root file system in order to edit following files for changing the node name:# mount /dev/dsk/c1t2d0s0 /mntChange the hostname from 0b to 0a:# cd /mnt/etc# vi hosts # vi hostname.bge0 # vi hostname.bge2 # vi nodename 4) Change the /mnt/etc/vfstab from the actual:/dev/md/dsk/d201        -       -       swap    -       no      -/dev/md/dsk/d200        /dev/md/rdsk/d200       /       ufs     1       no      -/dev/md/dsk/d205        /dev/md/rdsk/d205       /usr    ufs     1       no      logging/dev/md/dsk/d204        /dev/md/rdsk/d204       /var    ufs     1       no      logging#/dev/md/dsk/d206       /dev/md/rdsk/d206       /globaldevices  ufs     2       yes     loggingswap    -       /tmp    tmpfs   -       yes     -/dev/md/dsk/d206        /dev/md/rdsk/d206       /global/.devices/node@2 ufs     2       noglobalto this (unencapsulate disk from SDS/SVM):/dev/dsk/c1t0d0s1        -       -       swap    -       no      -/dev/dsk/c1t0d0s0       /dev/rdsk/c1t0d0s0       /       ufs     1       no      -/dev/dsk/c1t0d0s5       /dev/rdsk/c1t0d0s5       /usr    ufs     1       no      logging/dev/dsk/c1t0d0s4       /dev/rdsk/c1t0d0s4       /var    ufs     1       no      logging#/dev/md/dsk/d206       /dev/md/rdsk/d206       /globaldevices  ufs     2       yes     loggingswap    -       /tmp    tmpfs   -       yes     -/dev/dsk/c1t0d0s6       /dev/rdsk/c1t0d0s6       /global/.devices/node@1 ufs     2       no globalIt is important that global device partition (slice 6) in the new vfstab will point to the physical partition of the disk (in our case slice 6).Be careful with the name you use for the new disk. In this case we define it as c1t0d0 because we will insert it as target 0 in node 0a.But this could be different based on the configuration you are working on.5) Remove following entry from /mnt/etc/system (part of unencapsulation procedure):rootdev:/pseudo/md@0:0,200,blk6) Correct the link shared -> ../../global/.devices/node@2/dev/md/shared in order to point to the nodeid of node 0a (in our case nodeid 1):# cd /mnt/dev/mdhow it is now.... node 0b has nodeid 2lrwxrwxrwx   1 root     root          42 Mar 10  2005 shared ->../../global/.devices/node@2/dev/md/shared# rm shared# ln -s ../../global/.devices/node@1/dev/md/shared sharedhow is going to be... with nodeid 1 for node 0alrwxrwxrwx   1 root     root          42 Mar 10  2005 shared ->../../global/.devices/node@1/dev/md/shared7) Change nodeid (in our case from 2 to 1):# cd /mnt/etc/cluster# vi nodeid8) Change the file /mnt/etc/path_to_inst in order to reflect the correct nodeid for node 0a:# cd /mnt/etc# vi path_to_instChange entries from node@2 to node@1 with the vi command ":%s/node@2/node@1/g"9) Write the bootblock to the disk... just in case:# /usr/sbin/installboot /usr/platform/sun4u/lib/fs/ufs/bootblk /dev/rdsk/c1t2d0s0Now the disk is ready to be inserted in node 0a in order to bootup the node.10) Bootup node 0a with command "boot -sx"... this is becasue we need to make some changes in ccr files in order to recreate did environment.11) Modify cluster ccr:# cd /etc/cluster/ccr# rm did_instances# rm did_instances.bak# vi directory - remove the did_instances line.# /usr/cluster/lib/sc/ccradm -i /etc/cluster/ccr/directory # grep ccr_gennum /etc/cluster/ccr/directory ccr_gennum -1 # /usr/cluster/lib/sc/ccradm -i /etc/cluster/ccr/infrastructure # grep ccr_gennum /etc/cluster/ccr/infrastructure ccr_gennum -112) Bring the node 0a down again to the ok prompt and then issue the command "boot -r"Now the node will join the cluster and from scstat and metaset command you can verify functionality. Next step is to encapsulate the boot disk in SDS/SVM and create the mirrors.In our case node 0b has metadevice name starting from d200. For this reason on node 0a we need to create metadevice starting from d100. This is just an example, you can have different names.The important thing to remember is that metadevice boot disks have different names on each node.13) Remove metadevice pointing to the boot and mirror disks (inherit from node 0b):# metaclear -r -f d200# metaclear -r -f d201# metaclear -r -f d204# metaclear -r -f d205# metaclear -r -f d206verify from metastat that no metadevices are set for boot and mirror disks.14) Encapsulate the boot disk:# metainit -f d110 1 1 c1t0d0s0# metainit d100 -m d110# metaroot d10015) Reboot node 0a.16) Create all the metadevice for slices remaining on boot disk# metainit -f d111 1 1 c1t0d0s1# metainit d101 -m d111# metainit -f d114 1 1 c1t0d0s4# metainit d104 -m d114# metainit -f d115 1 1 c1t0d0s5# metainit d105 -m d115# metainit -f d116 1 1 c1t0d0s6# metainit d106 -m d11617) Edit the vfstab in order to specifiy metadevices created:old:/dev/dsk/c1t0d0s1        -       -       swap    -       no      -/dev/md/dsk/d100        /dev/md/rdsk/d100       /       ufs     1       no      -/dev/dsk/c1t0d0s5       /dev/rdsk/c1t0d0s5       /usr    ufs     1       no      logging/dev/dsk/c1t0d0s4       /dev/rdsk/c1t0d0s4       /var    ufs     1       no      logging#/dev/md/dsk/d206       /dev/md/rdsk/d206       /globaldevices  ufs     2       yes     loggingswap    -       /tmp    tmpfs   -       yes     -/dev/dsk/c1t0d0s6       /dev/rdsk/c1t0d0s6       /global/.devices/node@1 ufs      2       no  globalnew:/dev/md/dsk/d101        -       -       swap    -       no      -/dev/md/dsk/d100        /dev/md/rdsk/d100       /       ufs     1       no      -/dev/md/dsk/d105        /dev/md/rdsk/d105       /usr    ufs     1       no      logging/dev/md/dsk/d104        /dev/md/rdsk/d104       /var    ufs     1       no      logging#/dev/md/dsk/106       /dev/md/rdsk/d106       /globaldevices  ufs     2       yes     loggingswap    -       /tmp    tmpfs   -       yes     -/dev/md/dsk/d106        /dev/md/rdsk/d106       /global/.devices/node@1 ufs     2       noglobal18) Reboot node 0a in order to check new SDS/SVM boot configuration.19) Label the mirror disk c1t1d0 with the VTOC of boot disk c1t0d0:# prtvtoc /dev/dsk/c1t0d0s2 > /var/tmp/VTOC_c1t0d0 # fmthard -s /var/tmp/VTOC_c1t0d0 /dev/rdsk/c1t1d0s220) Put DB replica on slice 7 of disk c1t1d0:# metadb -a -c 3 /dev/dsk/c1t1d0s721) Create metadevice for mirror disk c1t1d0 and attach the new mirror side:# metainit d120 1 1 c1t1d0s0# metattach d100 d120# metainit d121 1 1 c1t1d0s1# metattach d101 d121# metainit d124 1 1 c1t1d0s4# metattach d104 d124# metainit d125 1 1 c1t1d0s5# metattach d105 d125# metainit d126 1 1 c1t1d0s6# metattach d106 d126

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  • Configure Jenkins and Tomcat using Puppet on Vagrant

    - by ex3v
    I'm playing with setting up my first Spring + jenkins + Tomcat CI dev environment. For now it's just a test/fun phase, but in the near future I'll be starting new project with my coworkers. That's the reason that I want development environment virtualized and exactly te same on every development machine, as well as on production server. I choosen to use Vagrant and to try to write puppet scripts that not only install everything, but also configure everything so each of us will have the same jenkins plugins, same jenkins and tomcat login and password, and literally after calling vagrant up we are ready to work. What I managed to do so far is installation of stuff needed and port forwarding. My vagrantfile looks like this (comments stripped): VAGRANTFILE_API_VERSION = "2" Vagrant.configure(VAGRANTFILE_API_VERSION) do |config| config.vm.box = "precise32" config.vm.box_url = "http://files.vagrantup.com/precise32.box" config.vm.network :forwarded_port, guest: 80, host: 8090 config.vm.network :forwarded_port, guest: 8080, host: 8091 config.vm.network :private_network, ip: "192.168.33.10" config.vm.provision :puppet do |puppet| puppet.manifests_path = "puppet/" puppet.manifest_file = "default.pp" puppet.options = ['--verbose'] end end And this is my puppet file: Exec { path => [ "/bin/", "/sbin/" , "/usr/bin/", "/usr/sbin/" ] } class system-update { exec { 'apt-get update': command => 'apt-get update', } $sysPackages = [ "build-essential" ] package { $sysPackages: ensure => "installed", require => Exec['apt-get update'], } } class tomcat { package { "tomcat": ensure => present, require => Class["system-update"], } service { "tomcat": ensure => "running", require => Package["tomcat"], } } class jenkins { package { "jenkins": ensure => present, require => Class["system-update"], } service { "jenkins": ensure => "running", require => Package["jenkins"], } } include system-update include tomcat include jenkins Now, when I hit vagrant provision and go to http://localhost:8091/ I can see jenkins running, so above script works good. Next step is configurating jenkins and tomcat by extending above puppet scripts. I'm pretty green when it comes to CI. After wandering around web I've found few tutorials about jenkins configuration (here's one of them). I really want to move configuration presented in this tutorial to puppet file, so when I spread my vagrantfile and puppet file between my coworkers, I will be sure that everyone has exactly te same setup. Unfortunately I'm also green about using puppet, I don't know how to do this. Any help will be apreciated.

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  • Connect to QuickBooks from PowerBuilder using RSSBus ADO.NET Data Provider

    - by dataintegration
    The RSSBus ADO.NET providers are easy-to-use, standards based controls that can be used from any platform or development technology that supports Microsoft .NET, including Sybase PowerBuilder. In this article we show how to use the RSSBus ADO.NET Provider for QuickBooks in PowerBuilder. A similar approach can be used from PowerBuilder with other RSSBus ADO.NET Data Providers to access data from Salesforce, SharePoint, Dynamics CRM, Google, OData, etc. In this article we will show how to create a basic PowerBuilder application that performs CRUD operations using the RSSBus ADO.NET Provider for QuickBooks. Step 1: Open PowerBuilder and create a new WPF Window Application solution. Step 2: Add all the Visual Controls needed for the connection properties. Step 3: Add the DataGrid control from the .NET controls. Step 4:Configure the columns of the DataGrid control as shown below. The column bindings will depend on the table. <DataGrid AutoGenerateColumns="False" Margin="13,249,12,14" Name="datagrid1" TabIndex="70" ItemsSource="{Binding}"> <DataGrid.Columns> <DataGridTextColumn x:Name="idColumn" Binding="{Binding Path=ID}" Header="ID" Width="SizeToHeader" /> <DataGridTextColumn x:Name="nameColumn" Binding="{Binding Path=Name}" Header="Name" Width="SizeToHeader" /> ... </DataGrid.Columns> </DataGrid> Step 5:Add a reference to the RSSBus ADO.NET Provider for QuickBooks assembly. Step 6:Optional: Set the QBXML Version to 6. Some of the tables in QuickBooks require a later version of QuickBooks to support updates and deletes. Please check the help for details. Connect the DataGrid: Once the visual elements have been configured, developers can use standard ADO.NET objects like Connection, Command, and DataAdapter to populate a DataTable with the results of a SQL query: System.Data.RSSBus.QuickBooks.QuickBooksConnection conn conn = create System.Data.RSSBus.QuickBooks.QuickBooksConnection(connectionString) System.Data.RSSBus.QuickBooks.QuickBooksCommand comm comm = create System.Data.RSSBus.QuickBooks.QuickBooksCommand(command, conn) System.Data.DataTable table table = create System.Data.DataTable System.Data.RSSBus.QuickBooks.QuickBooksDataAdapter dataAdapter dataAdapter = create System.Data.RSSBus.QuickBooks.QuickBooksDataAdapter(comm) dataAdapter.Fill(table) datagrid1.ItemsSource=table.DefaultView The code above can be used to bind data from any query (set this in command), to the DataGrid. The DataGrid should have the same columns as those returned from the SELECT statement. PowerBuilder Sample Project The included sample project includes the steps outlined in this article. You will also need the QuickBooks ADO.NET Data Provider to make the connection. You can download a free trial here.

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  • Using lookahead assertions in regular expressions

    - by Greg Jackson
    I use regular expressions on a daily basis, as my daily work is 90% in Perl (legacy codebase, but that's a different issue). Despite this, I still find lookahead and lookbehind to be terribly confusing and often unreadable. Right now, if I were to get a code review with a lookahead or lookbehind, I would immediately send it back to see if the problem can be solved by using multiple regular expressions or a different approach. The following are the main reasons I tend not to like them: They can be terribly unreadable. Lookahead assertions, for example, start from the beginning of the string no matter where they are placed. That, among other things, can cause some very "interesting" and non-obvious behaviors. It used to be the case that many languages didn't support lookahead/lookbehind (or supported them as "experimental features"). This isn't the case quite as much, but there's still always the question as to how well it's supported. Quite frankly, they feel like a dirty hack. Regexps often already are, but they can also be quite elegant, and have gained widespread acceptance. I've gotten by without any need for them at all... sometimes I think that they're extraneous. Now, I'll freely admit that especially the last two reasons aren't really good ones, but I felt that I should enumerate what goes through my mind when I see one. I'm more than willing to change my mind about them, but I feel that they violate some of my core tenets of programming, including: Code should be as readable as possible without sacrificing functionality -- this may include doing something in a less efficient, but clearer was as long as the difference is negligible or unimportant to the application as a whole. Code should be maintainable -- if another programmer comes along to fix my code, non-obvious behavior can hide bugs or make functional code appear buggy (see readability) "The right tool for the right job" -- I'm sure you can come up with contrived examples that could use lookahead, but I've never come across something that really needs them in my real-world development work. Is there anything that they're really the best tool for, as opposed to, say, multiple regexps (or, alternatively, are they the best tool for most cases they're used for today). My question is this: Is it good practice to use lookahead/lookbehind in regular expressions, or are they simply a hack that have found their way into modern production code? I'd be perfectly happy to be convinced that I'm wrong about this, and simple examples are useful for examples or illustration, but by themselves, won't be enough to convince me.

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  • Export all SSIS packages from msdb using Powershell

    - by jamiet
    Have you ever wanted to dump all the SSIS packages stored in msdb out to files? Of course you have, who wouldn’t? Right? Well, at least one person does because this was the subject of a thread (save all ssis packages to file) on the SSIS forum earlier today. Some of you may have already figured out a way of doing this but for those that haven’t here is a nifty little script that will do it for you and it uses our favourite jack-of-all tools … Powershell!! Imagine I have the following package folder structure on my Integration Services server (i.e. in [msdb]): There are two packages in there called “20110111 Chaining Expression components” & “Package”, I want to export those two packages into a folder structure that mirrors that in [msdb]. Here is the Powershell script that will do that:Param($SQLInstance = "localhost") #####Add all the SQL goodies (including Invoke-Sqlcmd)##### add-pssnapin sqlserverprovidersnapin100 -ErrorAction SilentlyContinue add-pssnapin sqlservercmdletsnapin100 -ErrorAction SilentlyContinue cls $Packages = Invoke-Sqlcmd -MaxCharLength 10000000 -ServerInstance $SQLInstance -Query "WITH cte AS ( SELECT cast(foldername as varchar(max)) as folderpath, folderid FROM msdb..sysssispackagefolders WHERE parentfolderid = '00000000-0000-0000-0000-000000000000' UNION ALL SELECT cast(c.folderpath + '\' + f.foldername as varchar(max)), f.folderid FROM msdb..sysssispackagefolders f INNER JOIN cte c ON c.folderid = f.parentfolderid ) SELECT c.folderpath,p.name,CAST(CAST(packagedata AS VARBINARY(MAX)) AS VARCHAR(MAX)) as pkg FROM cte c INNER JOIN msdb..sysssispackages p ON c.folderid = p.folderid WHERE c.folderpath NOT LIKE 'Data Collector%'" Foreach ($pkg in $Packages) { $pkgName = $Pkg.name $folderPath = $Pkg.folderpath $fullfolderPath = "c:\temp\$folderPath\" if(!(test-path -path $fullfolderPath)) { mkdir $fullfolderPath | Out-Null } $pkg.pkg | Out-File -Force -encoding ascii -FilePath "$fullfolderPath\$pkgName.dtsx" } To run it simply change the “localhost” parameter of the server you want to connect to either by editing the script or passing it in when the script is executed. It will create the folder structure in C:\Temp (which you can also easily change if you so wish – just edit the script accordingly). Here’s the folder structure that it created for me: Notice how it is a mirror of the folder structure in [msdb]. Hope this is useful! @Jamiet UPDATE: THis post prompted Chad Miller to write a post describing his Powershell add-in that utilises a SSIS API to do exporting of packages. Go take a read here: http://sev17.com/2011/02/importing-and-exporting-ssis-packages-using-powershell/

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  • Using Movemail with Thunderbird on Ubuntu

    - by rxt
    I'm trying to read local mail with Thunderbird on Ubuntu (with both 12.04 and 13.04). I've followed the instructions found here: How can I access system mail in /var/mail/ via thunderbird? I can read mail on the system using alpine or vim, so I know the mailbox is not empty. When I click the get-mail button, nothing happens. I see no Inbox (or any folder structure) for the specific account. I've set the rights for /var/mail to 1777. Settings server name: localhost username: john How can I get this working? Okay, considering the extra bounty, I would like to get this working like normal mail. The accepted answer from Qasim resulted in a much more usable situation than before - opening mail in Thunderbird with layout. I still face three problems though. When new mail is received in the mailbox, Thunderbird won't see this until after I restart Thunderbird. When Thunderbird is restarted, all mail is reset to unread and deleted mail is undone. This is probably because Thunderbird reads the mail from the /var/mail/www-data file, but doesn't update this file. So after restarting, it simply reads this file again, with the new mail and all old mail. This is probably a postfix issue: mail is sent out to existing mail addresses, but cannot be delivered because the receiving mailserver cannot be reached. This results in "Undelivered mail returned to sender". Only one mailserver can be reached: localhost. Because this is a test system, I don't want real customers to receive mail. I've blocked mail ports in UFW to be sure. When opening the returned mail, I can scroll down and then I see the original mail with proper layout. So I can read the mail, see if the proper images are included, and for me that's workable. Having to restart TB to read new mail - I know when new mail arrives, so I know when to restart. Having old mail restored after a restart - not big problem as well. I can delete the mail file if it gets too much. I know how it works, but it would be nice if it worked like normal.

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  • How to correctly track the analytics when using iframe

    - by Sherry Ann Hernandez
    In our main aspx page we have this analytics code <script type="text/javascript"> var _gaq = _gaq || []; _gaq.push(['_setAccount', 'UA-1301114-2']); _gaq.push(['_setDomainName', 'florahospitality.com']); _gaq.push(['_setAllowLinker', true]); _gaq.push(['_trackPageview']); _gaq.push(function() { var pageTracker = _gat._getTrackerByName(); var iframe = document.getElementById('reservationFrame'); iframe.src = pageTracker._getLinkerUrl('https://reservations.synxis.com/xbe/rez.aspx?Hotel=15159&template=flex&shell=flex&Chain=5375&locale=en&arrive=11/12/2012&depart=11/13/2012&adult=2&child=0&rooms=1&start=availresults&iata=&promo=&group='); }); (function() { var ga = document.createElement('script'); ga.type = 'text/javascript'; ga.async = true; ga.src = ('https:' == document.location.protocol ? 'https://ssl' : 'http://www') + '.google-analytics.com/ga.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(ga, s); })(); </script> Then inside this aspx page is an iframe. Inside the iframe we setup this analytics code <script type="text/javascript"> var _gaq = _gaq || []; _gaq.push(['_setAccount', 'UA-1301114-2']); _gaq.push(['_setDomainName', 'reservations.synxis.com']); _gaq.push(['_setAllowLinker', true]); _gaq.push(['_trackPageview', 'AvailabilityResults']); (function() { var ga = document.createElement('script'); ga.type = 'text/javascript'; ga.async = true; ga.src = ('https:' == document.location.protocol ? 'https://ssl' : 'http://www') + '.google-analytics.com/ga.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(ga, s); })(); </script> The problem is I see to pageview when I go to find the AvailabilityResults page. The first one is a direct traffic and the other one is a cpc. How come that they have different source? I was expecting that both of them is using a direct traffic.

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  • Automate RAC Cluster Upgrades using EM12c

    - by HariSrinivasan
    One of the most arduous processes  in DB maintenance is upgrading Databases across major versions, especially for complex RAC Clusters.With the release of Database Plug-in  (12.1.0.5.0), EM12c Rel 3 (12.1.0.3.0)  now supports automated upgrading of RAC Clusters in addition to Standalone Databases. This automation includes: Upgrade of the complete Cluster across the nodes. ( Example: 11.1.0.7 CRS, ASM, RAC DB  ->   11.2.0.4 or 12.1.0.1 GI, RAC DB)  Best practices in tune with your operations, where you can automate upgrade in steps: Step 1: Upgrade the Clusterware to Grid Infrastructure (Allowing you to wait, test and then move to DBs). Step 2: Upgrade RAC DBs either separately or in group (Mass upgrade of RAC DB's in the cluster). Standard pre-requisite checks like Cluster Verification Utility (CVU) and RAC checks Division of Upgrade process into Non-downtime activities (like laying down the new Oracle Homes (OH), running checks) to Downtime Activities (like Upgrading Clusterware to GI, Upgrading RAC) there by lowering the downtime required. Ability to configure Back up and Restore options as a part of this upgrade process. You can choose to : a. Take Backup via this process (either Guaranteed Restore Point (GRP) or RMAN) b. Set the procedure to pause just before the upgrade step to allow you to take a custom backup c. Ignore backup completely, if there are external mechanisms already in place.  High Level Steps: Select the Procedure "Upgrade Database" from Database Provisioning Home page. Choose the Target Type for upgrade and the Destination version Pick and choose the Cluster, it picks up the complete topology since the clusterware/GI isn't upgraded already Select the Gold Image of the destination version for deploying both the GI and RAC OHs Specify new OH patch, credentials, choose the Restore and Backup options, if required provide additional pre and post scripts Set the Break points in the procedure execution to isolate Downtime activities Submit and track the procedure's execution status.  The animation below captures the steps in the wizard.  For step by step process and to understand the support matrix check this documentation link. Explore the functionality!! In the next blog, will talk about automating rolling Upgrades of Databases in Physical Standby Data Guard environment using Transient Logical Standby.

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  • Render To Texture Using OpenGL is not working but normal rendering works just fine

    - by Franky Rivera
    things I initialize at the beginning of the program I realize not all of these pertain to my issue I just copy and pasted what I had //overall initialized //things openGL related I initialize earlier on in the project glClearColor( 0.0f, 0.0f, 0.0f, 1.0f ); glClearDepth( 1.0f ); glEnable(GL_ALPHA_TEST); glEnable( GL_STENCIL_TEST ); glEnable(GL_DEPTH_TEST); glDepthFunc( GL_LEQUAL ); glEnable(GL_CULL_FACE); glFrontFace( GL_CCW ); glEnable(GL_COLOR_MATERIAL); glEnable(GL_BLEND); glBlendFunc(GL_SRC_ALPHA, GL_ONE_MINUS_SRC_ALPHA); glHint( GL_PERSPECTIVE_CORRECTION_HINT, GL_NICEST ); //we also initialize our shader programs //(i added some shader program functions for definitions) //this enum list is else where in code //i figured it would help show you guys more about my //shader compile creation function right under this enum list VVVVVV /*enum eSHADER_ATTRIB_LOCATION { VERTEX_ATTRIB = 0, NORMAL_ATTRIB = 2, COLOR_ATTRIB, COLOR2_ATTRIB, FOG_COORD, TEXTURE_COORD_ATTRIB0 = 8, TEXTURE_COORD_ATTRIB1, TEXTURE_COORD_ATTRIB2, TEXTURE_COORD_ATTRIB3, TEXTURE_COORD_ATTRIB4, TEXTURE_COORD_ATTRIB5, TEXTURE_COORD_ATTRIB6, TEXTURE_COORD_ATTRIB7 }; */ //if we fail making our shader leave if( !testShader.CreateShader( "SimpleShader.vp", "SimpleShader.fp", 3, VERTEX_ATTRIB, "vVertexPos", NORMAL_ATTRIB, "vNormal", TEXTURE_COORD_ATTRIB0, "vTexCoord" ) ) return false; if( !testScreenShader.CreateShader( "ScreenShader.vp", "ScreenShader.fp", 3, VERTEX_ATTRIB, "vVertexPos", NORMAL_ATTRIB, "vNormal", TEXTURE_COORD_ATTRIB0, "vTexCoord" ) ) return false; SHADER PROGRAM FUNCTIONS bool CShaderProgram::CreateShader( const char* szVertexShaderName, const char* szFragmentShaderName, ... ) { //here are our handles for the openGL shaders int iGLVertexShaderHandle = -1, iGLFragmentShaderHandle = -1; //get our shader data char *vData = 0, *fData = 0; int vLength = 0, fLength = 0; LoadShaderFile( szVertexShaderName, &vData, &vLength ); LoadShaderFile( szFragmentShaderName, &fData, &fLength ); //data if( !vData ) return false; //data if( !fData ) { delete[] vData; return false; } //create both our shader objects iGLVertexShaderHandle = glCreateShader( GL_VERTEX_SHADER ); iGLFragmentShaderHandle = glCreateShader( GL_FRAGMENT_SHADER ); //well we got this far so we have dynamic data to clean up //load vertex shader glShaderSource( iGLVertexShaderHandle, 1, (const char**)(&vData), &vLength ); //load fragment shader glShaderSource( iGLFragmentShaderHandle, 1, (const char**)(&fData), &fLength ); //we are done with our data delete it delete[] vData; delete[] fData; //compile them both glCompileShader( iGLVertexShaderHandle ); //get shader status int iShaderOk; glGetShaderiv( iGLVertexShaderHandle, GL_COMPILE_STATUS, &iShaderOk ); if( iShaderOk == GL_FALSE ) { char* buffer; //get what happend with our shader glGetShaderiv( iGLVertexShaderHandle, GL_INFO_LOG_LENGTH, &iShaderOk ); buffer = new char[iShaderOk]; glGetShaderInfoLog( iGLVertexShaderHandle, iShaderOk, NULL, buffer ); //sprintf_s( buffer, "Failure Our Object For %s was not created", szFileName ); MessageBoxA( NULL, buffer, szVertexShaderName, MB_OK ); //delete our dynamic data free( buffer ); glDeleteShader(iGLVertexShaderHandle); return false; } glCompileShader( iGLFragmentShaderHandle ); //get shader status glGetShaderiv( iGLFragmentShaderHandle, GL_COMPILE_STATUS, &iShaderOk ); if( iShaderOk == GL_FALSE ) { char* buffer; //get what happend with our shader glGetShaderiv( iGLFragmentShaderHandle, GL_INFO_LOG_LENGTH, &iShaderOk ); buffer = new char[iShaderOk]; glGetShaderInfoLog( iGLFragmentShaderHandle, iShaderOk, NULL, buffer ); //sprintf_s( buffer, "Failure Our Object For %s was not created", szFileName ); MessageBoxA( NULL, buffer, szFragmentShaderName, MB_OK ); //delete our dynamic data free( buffer ); glDeleteShader(iGLFragmentShaderHandle); return false; } //lets check to see if the fragment shader compiled int iCompiled = 0; glGetShaderiv( iGLVertexShaderHandle, GL_COMPILE_STATUS, &iCompiled ); if( !iCompiled ) { //this shader did not compile leave return false; } //lets check to see if the fragment shader compiled glGetShaderiv( iGLFragmentShaderHandle, GL_COMPILE_STATUS, &iCompiled ); if( !iCompiled ) { char* buffer; //get what happend with our shader glGetShaderiv( iGLFragmentShaderHandle, GL_INFO_LOG_LENGTH, &iShaderOk ); buffer = new char[iShaderOk]; glGetShaderInfoLog( iGLFragmentShaderHandle, iShaderOk, NULL, buffer ); //sprintf_s( buffer, "Failure Our Object For %s was not created", szFileName ); MessageBoxA( NULL, buffer, szFragmentShaderName, MB_OK ); //delete our dynamic data free( buffer ); glDeleteShader(iGLFragmentShaderHandle); return false; } //make our new shader program m_iShaderProgramHandle = glCreateProgram(); glAttachShader( m_iShaderProgramHandle, iGLVertexShaderHandle ); glAttachShader( m_iShaderProgramHandle, iGLFragmentShaderHandle ); glLinkProgram( m_iShaderProgramHandle ); int iLinked = 0; glGetProgramiv( m_iShaderProgramHandle, GL_LINK_STATUS, &iLinked ); if( !iLinked ) { //we didn't link return false; } //NOW LETS CREATE ALL OUR HANDLES TO OUR PROPER LIKING //start from this parameter va_list parseList; va_start( parseList, szFragmentShaderName ); //read in number of variables if any unsigned uiNum = 0; uiNum = va_arg( parseList, unsigned ); //for loop through our attribute pairs int enumType = 0; for( unsigned x = 0; x < uiNum; ++x ) { //specify our attribute locations enumType = va_arg( parseList, int ); char* name = va_arg( parseList, char* ); glBindAttribLocation( m_iShaderProgramHandle, enumType, name ); } //end our list parsing va_end( parseList ); //relink specify //we have custom specified our attribute locations glLinkProgram( m_iShaderProgramHandle ); //fill our handles InitializeHandles( ); //everything went great return true; } void CShaderProgram::InitializeHandles( void ) { m_uihMVP = glGetUniformLocation( m_iShaderProgramHandle, "mMVP" ); m_uihWorld = glGetUniformLocation( m_iShaderProgramHandle, "mWorld" ); m_uihView = glGetUniformLocation( m_iShaderProgramHandle, "mView" ); m_uihProjection = glGetUniformLocation( m_iShaderProgramHandle, "mProjection" ); ///////////////////////////////////////////////////////////////////////////////// //texture handles m_uihDiffuseMap = glGetUniformLocation( m_iShaderProgramHandle, "diffuseMap" ); if( m_uihDiffuseMap != -1 ) { //store what texture index this handle will be in the shader glUniform1i( m_uihDiffuseMap, RM_DIFFUSE+GL_TEXTURE0 ); (0)+ } m_uihNormalMap = glGetUniformLocation( m_iShaderProgramHandle, "normalMap" ); if( m_uihNormalMap != -1 ) { //store what texture index this handle will be in the shader glUniform1i( m_uihNormalMap, RM_NORMAL+GL_TEXTURE0 ); (1)+ } } void CShaderProgram::SetDiffuseMap( const unsigned& uihDiffuseMap ) { (0)+ glActiveTexture( RM_DIFFUSE+GL_TEXTURE0 ); glBindTexture( GL_TEXTURE_2D, uihDiffuseMap ); } void CShaderProgram::SetNormalMap( const unsigned& uihNormalMap ) { (1)+ glActiveTexture( RM_NORMAL+GL_TEXTURE0 ); glBindTexture( GL_TEXTURE_2D, uihNormalMap ); } //MY 2 TEST SHADERS also my math order is correct it pertains to my matrix ordering in my math library once again i've tested the basic rendering. rendering to the screen works fine ----------------------------------------SIMPLE SHADER------------------------------------- //vertex shader looks like this #version 330 in vec3 vVertexPos; in vec3 vNormal; in vec2 vTexCoord; uniform mat4 mWorld; // Model Matrix uniform mat4 mView; // Camera View Matrix uniform mat4 mProjection;// Camera Projection Matrix out vec2 vTexCoordVary; // Texture coord to the fragment program out vec3 vNormalColor; void main( void ) { //pass the texture coordinate vTexCoordVary = vTexCoord; vNormalColor = vNormal; //calculate our model view projection matrix mat4 mMVP = (( mWorld * mView ) * mProjection ); //result our position gl_Position = vec4( vVertexPos, 1 ) * mMVP; } //fragment shader looks like this #version 330 in vec2 vTexCoordVary; in vec3 vNormalColor; uniform sampler2D diffuseMap; uniform sampler2D normalMap; out vec4 fragColor[2]; void main( void ) { //CORRECT fragColor[0] = texture( normalMap, vTexCoordVary ); fragColor[1] = vec4( vNormalColor, 1.0 ); }; ----------------------------------------SCREEN SHADER------------------------------------- //vertext shader looks like this #version 330 in vec3 vVertexPos; // This is the position of the vertex coming in in vec2 vTexCoord; // This is the texture coordinate.... out vec2 vTexCoordVary; // Texture coord to the fragment program void main( void ) { vTexCoordVary = vTexCoord; //set our position gl_Position = vec4( vVertexPos.xyz, 1.0f ); } //fragment shader looks like this #version 330 in vec2 vTexCoordVary; // Incoming "varying" texture coordinate uniform sampler2D diffuseMap;//the tile detail texture uniform sampler2D normalMap; //the normal map from earlier out vec4 vTheColorOfThePixel; void main( void ) { //CORRECT vTheColorOfThePixel = texture( normalMap, vTexCoordVary ); }; .Class RenderTarget Main Functions //here is my render targets create function bool CRenderTarget::Create( const unsigned uiNumTextures, unsigned uiWidth, unsigned uiHeight, int iInternalFormat, bool bDepthWanted ) { if( uiNumTextures <= 0 ) return false; //generate our variables glGenFramebuffers(1, &m_uifboHandle); // Initialize FBO glBindFramebuffer(GL_FRAMEBUFFER, m_uifboHandle); m_uiNumTextures = uiNumTextures; if( bDepthWanted ) m_uiNumTextures += 1; m_uiTextureHandle = new unsigned int[uiNumTextures]; glGenTextures( uiNumTextures, m_uiTextureHandle ); for( unsigned x = 0; x < uiNumTextures-1; ++x ) { glBindTexture( GL_TEXTURE_2D, m_uiTextureHandle[x]); // Reserve space for our 2D render target glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MIN_FILTER, GL_LINEAR); glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MAG_FILTER, GL_LINEAR); glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_S, GL_CLAMP_TO_EDGE); glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_T, GL_CLAMP_TO_EDGE); glTexImage2D(GL_TEXTURE_2D, 0, iInternalFormat, uiWidth, uiHeight, 0, GL_RGB, GL_UNSIGNED_BYTE, NULL); glFramebufferTexture2D(GL_DRAW_FRAMEBUFFER, GL_COLOR_ATTACHMENT0 + x, GL_TEXTURE_2D, m_uiTextureHandle[x], 0); } //if we need one for depth testing if( bDepthWanted ) { glFramebufferTexture2D(GL_FRAMEBUFFER_EXT, GL_DEPTH_ATTACHMENT, GL_TEXTURE_2D, m_uiTextureHandle[uiNumTextures-1], 0); glFramebufferTexture2D(GL_FRAMEBUFFER_EXT, GL_STENCIL_ATTACHMENT, GL_TEXTURE_2D, m_uiTextureHandle[uiNumTextures-1], 0);*/ // Must attach texture to framebuffer. Has Stencil and depth glBindRenderbuffer(GL_RENDERBUFFER, m_uiTextureHandle[uiNumTextures-1]); glRenderbufferStorage(GL_RENDERBUFFER, /*GL_DEPTH_STENCIL*/GL_DEPTH24_STENCIL8, TEXTURE_WIDTH, TEXTURE_HEIGHT ); glFramebufferRenderbuffer(GL_FRAMEBUFFER, GL_DEPTH_ATTACHMENT, GL_RENDERBUFFER, m_uiTextureHandle[uiNumTextures-1]); glFramebufferRenderbuffer(GL_FRAMEBUFFER, GL_STENCIL_ATTACHMENT, GL_RENDERBUFFER, m_uiTextureHandle[uiNumTextures-1]); } glBindFramebuffer(GL_FRAMEBUFFER, 0); //everything went fine return true; } void CRenderTarget::Bind( const int& iTargetAttachmentLoc, const unsigned& uiWhichTexture, const bool bBindFrameBuffer ) { if( bBindFrameBuffer ) glBindFramebuffer( GL_FRAMEBUFFER, m_uifboHandle ); if( uiWhichTexture < m_uiNumTextures ) glFramebufferTexture(GL_FRAMEBUFFER, GL_COLOR_ATTACHMENT0 + iTargetAttachmentLoc, m_uiTextureHandle[uiWhichTexture], 0); } void CRenderTarget::UnBind( void ) { //default our binding glBindFramebuffer( GL_FRAMEBUFFER, 0 ); } //this is all in a test project so here's my straight forward rendering function for testing this render function does basic rendering steps keep in mind i have already tested my textures i have already tested my box thats being rendered all basic rendering works fine its just when i try to render to a texture then display it in a render surface that it does not work. Also I have tested my render surface it is bound exactly to the screen coordinate space void TestRenderSteps( void ) { //Clear the color and the depth glClearColor( 0.0f, 0.0f, 0.0f, 1.0f ); glClear( GL_COLOR_BUFFER_BIT | GL_DEPTH_BUFFER_BIT ); //bind the shader program glUseProgram( testShader.m_iShaderProgramHandle ); //1) grab the vertex buffer related to our rendering glBindBuffer( GL_ARRAY_BUFFER, CVertexBufferManager::GetInstance()->GetPositionNormalTexBuffer().GetBufferHandle() ); //2) how our stream will be split here ( 4 bytes position, ..ext ) CVertexBufferManager::GetInstance()->GetPositionNormalTexBuffer().MapVertexStride(); //3) set the index buffer if needed glBindBuffer( GL_ELEMENT_ARRAY_BUFFER, CIndexBuffer::GetInstance()->GetBufferHandle() ); //send the needed information into the shader testShader.SetWorldMatrix( boxPosition ); testShader.SetViewMatrix( Static_Camera.GetView( ) ); testShader.SetProjectionMatrix( Static_Camera.GetProjection( ) ); testShader.SetDiffuseMap( iTextureID ); testShader.SetNormalMap( iTextureID2 ); GLenum buffers[] = { GL_COLOR_ATTACHMENT0, GL_COLOR_ATTACHMENT1 }; glDrawBuffers(2, buffers); //bind to our render target //RM_DIFFUSE, RM_NORMAL are enums (0 && 1) renderTarget.Bind( RM_DIFFUSE, 1, true ); renderTarget.Bind( RM_NORMAL, 1, false); //false because buffer is already bound //i clear here just to clear the texture to make it a default value of white //by doing this i can see if what im rendering to my screen is just drawing to the screen //or if its my render target defaulted glClearColor( 1.0f, 1.0f, 1.0f, 1.0f ); glClear( GL_COLOR_BUFFER_BIT | GL_DEPTH_BUFFER_BIT ); //i have this box object which i draw testBox.Draw(); //the draw call looks like this //my normal rendering works just fine so i know this draw is fine // glDrawElementsBaseVertex( m_sides[x].GetPrimitiveType(), // m_sides[x].GetPrimitiveCount() * 3, // GL_UNSIGNED_INT, // BUFFER_OFFSET(sizeof(unsigned int) * m_sides[x].GetStartIndex()), // m_sides[x].GetStartVertex( ) ); //we unbind the target back to default renderTarget.UnBind(); //i stop mapping my vertex format CVertexBufferManager::GetInstance()->GetPositionNormalTexBuffer().UnMapVertexStride(); //i go back to default in using no shader program glUseProgram( 0 ); //now that everything is drawn to the textures //lets draw our screen surface and pass it our 2 filled out textures //NOW RENDER THE TEXTURES WE COLLECTED TO THE SCREEN QUAD //bind the shader program glUseProgram( testScreenShader.m_iShaderProgramHandle ); //1) grab the vertex buffer related to our rendering glBindBuffer( GL_ARRAY_BUFFER, CVertexBufferManager::GetInstance()->GetPositionTexBuffer().GetBufferHandle() ); //2) how our stream will be split here CVertexBufferManager::GetInstance()->GetPositionTexBuffer().MapVertexStride(); //3) set the index buffer if needed glBindBuffer( GL_ELEMENT_ARRAY_BUFFER, CIndexBuffer::GetInstance()->GetBufferHandle() ); //pass our 2 filled out textures (in the shader im just using the diffuse //i wanted to see if i was rendering anything before i started getting into other techniques testScreenShader.SetDiffuseMap( renderTarget.GetTextureHandle(0) ); //SetDiffuseMap definitions in shader program class testScreenShader.SetNormalMap( renderTarget.GetTextureHandle(1) ); //SetNormalMap definitions in shader program class //DO the draw call drawing our screen rectangle glDrawElementsBaseVertex( m_ScreenRect.GetPrimitiveType(), m_ScreenRect.GetPrimitiveCount() * 3, GL_UNSIGNED_INT, BUFFER_OFFSET(sizeof(unsigned int) * m_ScreenRect.GetStartIndex()), m_ScreenRect.GetStartVertex( ) );*/ //unbind our vertex mapping CVertexBufferManager::GetInstance()->GetPositionTexBuffer().UnMapVertexStride(); //default to no shader program glUseProgram( 0 ); } Last words: 1) I can render my box just fine 2) i can render my screen rect just fine 3) I cannot render my box into a texture then display it into my screen rect 4) This entire project is just a test project I made to test different rendering practices. So excuse any "ugly-ish" unclean code. This was made just on a fly run through when I was trying new test cases.

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  • Using a SQL Prompt snippet with template parameters

    - by SQLDev
    As part of my product management role I regularly attend trade shows and man the Red Gate booth in the vendor exhibition hall. Amongst other things this involves giving product demos to customers. Our latest demo involves SQL Source Control and SQL Test in a continuous integration environment. In order to demonstrate quite how easy it is to set up our tools from scratch we start the demo by creating an entirely new database to link to source control, using an individual database name for each conference attendee. In SQL Server Management Studio this can be done either by selecting New Database from the Object Explorer or by executing “CREATE DATABASE DemoDB_John” in a query window. We recently extended the demo to include SQL Test. This uses an open source SQL Server unit testing framework called tSQLt (www.tsqlt.org), which has a CLR object that requires EXTERNAL_ACCESS to be set as follows: ALTER DATABASE DemoDB_John SET TRUSTWORTHY ON This isn’t hard to do, but if you’re giving demo after demo, this two-step process soon becomes tedious. This is where SQL Prompt snippets come into their own. I can create a snippet named create_demo_db for this following: CREATE DATABASE DemoDB_John GO USE DemoDB_John GO ALTER DATABASE DemoDB_John SET TRUSTWORTHY ON Now I just have to type the first few characters of the snippet name, select the snippet from SQL Prompt’s candidate list, and execute the code. Simple! The problem is that this can only work once due to the hard-coded database name. Luckily I can leverage a nice feature in SQL Server Management Studio called Template Parameters. If I modify my snippet to be: CREATE DATABASE <DBName,, DemoDB_> GO USE <DBName,, DemoDB_> GO ALTER DATABASE <DBName,, DemoDB_> SET TRUSTWORTHY ON Once I’ve invoked the snippet, I can press Ctrl-Shift-M, which calls up the Specify Values for Template Parameters dialog, where I can type in my database name just once. Now you can click OK and run the query. Easy. Ideally I’d like for SQL Prompt to auto-invoke the Template Parameter dialog for all snippets where it detects the angled bracket syntax, but typing in the keyboard shortcut is a small price to pay for the time savings.

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  • Fraud and Anomaly Detection using Oracle Data Mining YouTube-like Video

    - by chberger
    I've created and recorded another YouTube-like presentation and "live" demos of Oracle Advanced Analytics Option, this time focusing on Fraud and Anomaly Detection using Oracle Data Mining.  [Note:  It is a large MP4 file that will open and play in place.  The sound quality is weak so you may need to turn up the volume.] Data is your most valuable asset. It represents the entire history of your organization and its interactions with your customers.  Predictive analytics leverages data to discover patterns, relationships and to help you even make informed predictions.   Oracle Data Mining (ODM) automatically discovers relationships hidden in data.  Predictive models and insights discovered with ODM address business problems such as:  predicting customer behavior, detecting fraud, analyzing market baskets, profiling and loyalty.  Oracle Data Mining, part of the Oracle Advanced Analytics (OAA) Option to the Oracle Database EE, embeds 12 high performance data mining algorithms in the SQL kernel of the Oracle Database. This eliminates data movement, delivers scalability and maintains security.  But, how do you find these very important needles or possibly fraudulent transactions and huge haystacks of data? Oracle Data Mining’s 1 Class Support Vector Machine algorithm is specifically designed to identify rare or anomalous records.  Oracle Data Mining's 1-Class SVM anomaly detection algorithm trains on what it believes to be considered “normal” records, build a descriptive and predictive model which can then be used to flags records that, on a multi-dimensional basis, appear to not fit in--or be different.  Combined with clustering techniques to sort transactions into more homogeneous sub-populations for more focused anomaly detection analysis and Oracle Business Intelligence, Enterprise Applications and/or real-time environments to "deploy" fraud detection, Oracle Data Mining delivers a powerful advanced analytical platform for solving important problems.  With OAA/ODM you can find suspicious expense report submissions, flag non-compliant tax submissions, fight fraud in healthcare claims and save huge amounts of money in fraudulent claims  and abuse.   This presentation and several brief demos will show Oracle Data Mining's fraud and anomaly detection capabilities.  

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  • Using jQuery to Make a Field Read Only in SharePoint

    - by Mark Rackley
    Okay… this will be my shortest blog post EVER. Very little rambling.. I promise, and I’m sure this has been blogged more than once, so I apologize for adding to the noise, but like I always say, I blog for myself so I have a global bookmark. So,let’s say you have a field on a SharePoint Form and you want to make it read only. You COULD just open it up in SPD and easily make it read only, but some people are purists and don’t like use SPD or modify the default new/edit/disp forms. Put me in the latter camp, I try to avoid modifying these forms and it seemed like such a simple task that I didn’t want to create a new un-ghosted form.  So.. how do you do it? It’s only one line of jQuery. All you need to do is find the id for your input field and capture the keypress on it so that it cannot be modified (you should probably capture clicks for dropdowns/checkboxes/etc. but I didn’t need to).  Anyway, here’s the entire script: <script src="jquery.min.js" type="text/javascript"></script> <script type="text/javascript"> jQuery(document).ready(function($){ //capture keypress on our read only field and return false $('#idOfInputField').keypress(function() { return false; }); }) </script>   You can find the ID of your input field by viewing the source, this ID stays consistent as long as you don’t muck with the list or form in the wrong way.  Please note, you CANNOT disable the input field as an alternative to capturing the keypress. If you do this and save the form, any data in the disabled fields will be wiped out. There are probably a dozen ways to make a field read-only and if all you are using jQuery for is to make a field read-only, then you might want to question your use of adding the overhead (although it’s really not that much). Hey.. it’s another tool for your tool belt.

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  • Upgrading 10.04LTS -> 10.10 using custom sources

    - by Boatzart
    I'm trying to upgrade to 10.10 from 10.04 LTS using a custom sources.list file that points to an unofficial mirror*. The mirror does have maverick, but I get the following output when upgrading: boatzart@somecomputer: > sudo do-release-upgrade Checking for a new ubuntu release Done Upgrade tool signature Done Upgrade tool Done downloading extracting 'maverick.tar.gz' authenticate 'maverick.tar.gz' against 'maverick.tar.gz.gpg' tar: Removing leading `/' from member names Reading cache Checking package manager Reading package lists... Done Building dependency tree Reading state information... Done Building data structures... Done Reading package lists... Done Building dependency tree Reading state information... Done Building data structures... Done Updating repository information WARNING: Failed to read mirror file No valid mirror found While scanning your repository information no mirror entry for the upgrade was found. This can happen if you run a internal mirror or if the mirror information is out of date. Do you want to rewrite your 'sources.list' file anyway? If you choose 'Yes' here it will update all 'lucid' to 'maverick' entries. If you select 'No' the upgrade will cancel. Continue [yN] y WARNING: Failed to read mirror file 96% [Working] Checking package manager Reading package lists... Done Building dependency tree Reading state information... Done Building data structures... Done Calculating the changes Calculating the changes Could not calculate the upgrade An unresolvable problem occurred while calculating the upgrade: The package 'update-manager-kde' is marked for removal but it is in the removal blacklist. This can be caused by: * Upgrading to a pre-release version of Ubuntu * Running the current pre-release version of Ubuntu * Unofficial software packages not provided by Ubuntu If none of this applies, then please report this bug against the 'update-manager' package and include the files in /var/log/dist-upgrade/ in the bug report. Restoring original system state Aborting Reading package lists... Done Building dependency tree Reading state information... Done Building data structures... Done Here is the relevant section from /var/log/dist-upgrade/main.log: 2010-11-18 14:05:52,117 DEBUG The package 'update-manager-kde' is marked for removal but it's in the removal blacklist 2010-11-18 14:05:52,136 ERROR Dist-upgrade failed: 'The package 'update-manager-kde' is marked for removal but it is in the removal blacklist.' 2010-11-18 14:05:52,136 DEBUG abort called *I'm located inside of USC, and for some crazy reason any sustained downloads to anywhere outside of the University are throttled down to 5kbps inside of my lab. Because of this I need to use the following sources.list: deb http://mirrors.usc.edu/pub/linux/distributions/ubuntu/ lucid main restricted universe multiverse deb http://mirrors.usc.edu/pub/linux/distributions/ubuntu/ lucid-updates main restricted universe multiverse deb http://mirrors.usc.edu/pub/linux/distributions/ubuntu/ lucid-backports main restricted universe multiverse deb http://mirrors.usc.edu/pub/linux/distributions/ubuntu/ lucid-security main restricted universe multiverse I've tried adding four more entries to the sources.list with s/lucid/maverick/ but that didn't help. Does anyone know how to fix this? Thanks!

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  • How to force a clock update using ntp?

    - by ysap
    I am running Ubuntu on an ARM based embedded system that lacks a battery backed RTC. The wake-up time is somewhere during 1970. Thus, I use the NTP service to update the time to the current time. I added the following line to /etc/rc.local file: sudo ntpdate -s time.nist.gov However, after startup, it still takes a couple of minutes until the time is updated, during which period I cannot work effectively with tar and make. How can I force a clock update at any given time? UPDATE 1: The following (thanks to Eric and Stephan) works fine from command line, but fails to update the clock when put in /etc/rc.local: $ date ; sudo service ntp stop ; sudo ntpdate -s time.nist.gov ; sudo service ntp start ; date Thu Jan 1 00:00:58 UTC 1970 * Stopping NTP server ntpd [ OK ] * Starting NTP server [ OK ] Thu Feb 14 18:52:21 UTC 2013 What am I doing wrong? UPDATE 2: I tried following the few suggestions that came in response to the 1st update, but nothing seems to actually do the job as required. Here's what I tried: Replace the server to us.pool.ntp.org Use explicit paths to the programs Remove the ntp service altogether and leave just sudo ntpdate ... in rc.local Remove the sudo from the above command in rc.local Using the above, the machine still starts at 1970. However, when doing this from command line once logged in (via ssh), the clock gets updated as soon as I invoke ntpdate. Last thing I did was to remove that from rc.local and place a call to ntpdate in my .bashrc file. This does update the clock as expected, and I get the true current time once the command prompt is available. However, this means that if the machine is turned on and no user is logged in, then the time never gets updates. I can, of course, reinstall the ntp service so at least the clock is updated within a few minutes from startup, but then we're back at square 1. So, is there a reason why placing the ntpdate command in rc.local does not perform the required task, while doing so in .bashrc works fine?

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  • Enabling EUS support in OUD 11gR2 using command line interface

    - by Sylvain Duloutre
    Enterprise User Security (EUS) allows Oracle Database to use users & roles stored in LDAP for authentication and authorization.Since the 11gR2 release, OUD natively supports EUS. EUS can be easily configured during OUD setup. ODSM (the graphical admin console) can also be used to enable EUS for a new suffix. However, enabling EUS for a new suffix using command line interface is currently not documented, so here is the procedure: Let's assume that EUS support was enabled during initial setup.Let's o=example be the new suffix I want to use to store Enterprise users. The following sequence of command must be applied for each new suffix: // Create a local database holding EUS context infodsconfig create-workflow-element --set base-dn:cn=OracleContext,o=example --set enabled:true --type db-local-backend --element-name exampleContext -n // Add a workflow element in the call path to generate on the fly attributes required by EUSdsconfig create-workflow-element --set enabled:true --type eus-context --element-name eusContext --set next-workflow-element:exampleContext -n // Add the context to a workflow for routingdsconfig create-workflow --set base-dn:cn=OracleContext,o=example --set enabled:true --set workflow-element:eusContext --workflow-name exampleContext_workflow -n //Add the new workflow to the appropriate network groupdsconfig set-network-group-prop --group-name network-group --add workflow:exampleContext_workflow -n // Create the local database for o=exampledsconfig create-workflow-element --set base-dn:o=example --set enabled:true --type db-local-backend --element-name example -n // Create a workflow element in the call path to the user data to generate on the fly attributes expected by EUS dsconfig create-workflow-element --set enabled:true --set eus-realm:o=example --set next-workflow-element:example --type eus --element-name eusWfe// Add the db to a workflow for routingdsconfig create-workflow --set base-dn:o=example --set enabled:true --set workflow-element:eusWfe --workflow-name example_workflow -n //Add the new workflow to the appropriate network groupdsconfig set-network-group-prop --group-name network-group --add workflow:example_workflow -n  // Add the appropriate acis for EUSdsconfig set-access-control-handler-prop \           --add global-aci:'(target="ldap:///o=example")(targetattr="authpassword")(version 3.0; acl "EUS reads authpassword"; allow (read,search,compare) userdn="ldap:///??sub?(&(objectclass=orclservice)(objectclass=orcldbserver))";)' dsconfig set-access-control-handler-prop \       --add global-aci:'(target="ldap:///o=example")(targetattr="orclaccountstatusevent")(version 3.0; acl "EUS writes orclaccountstatusenabled"; allow (write) userdn="ldap:///??sub?(&(objectclass=orclservice)(objectclass=orcldbserver))";)' Last but not least you must adapt the content of the ${OUD}/config/EUS/eusData.ldif  file with your suffix value then inport it into OUD.

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  • Using elapsed time for SlowMo in XNA

    - by Dave Voyles
    I'm trying to create a slow-mo effect in my pong game so that when a player is a button the paddles and ball will suddenly move at a far slower speed. I believe my understanding of the concepts of adjusting the timing in XNA are done, but I'm not sure of how to incorporate it into my design exactly. The updates for my bats (paddles) are done in my Bat.cs class: /// Controls the bat moving up the screen /// </summary> public void MoveUp() { SetPosition(Position + new Vector2(0, -moveSpeed)); } /// <summary> /// Controls the bat moving down the screen /// </summary> public void MoveDown() { SetPosition(Position + new Vector2(0, moveSpeed)); } /// <summary> /// Updates the position of the AI bat, in order to track the ball /// </summary> /// <param name="ball"></param> public virtual void UpdatePosition(Ball ball) { size.X = (int)Position.X; size.Y = (int)Position.Y; } While the rest of my game updates are done in my GameplayScreen.cs class (I'm using the XNA game state management sample) Class GameplayScreen { ........... bool slow; .......... public override void Update(GameTime gameTime, bool otherScreenHasFocus, bool coveredByOtherScreen) base.Update(gameTime, otherScreenHasFocus, false); if (IsActive) { // SlowMo Stuff Elapsed = (float)gameTime.ElapsedGameTime.TotalSeconds; if (Slowmo) Elapsed *= .8f; MoveTimer += Elapsed; double elapsedTime = gameTime.ElapsedGameTime.TotalMilliseconds; if (Keyboard.GetState().IsKeyDown(Keys.Up)) slow = true; else if (Keyboard.GetState().IsKeyDown(Keys.Down)) slow = false; if (slow == true) elapsedTime *= .1f; // Updating bat position leftBat.UpdatePosition(ball); rightBat.UpdatePosition(ball); // Updating the ball position ball.UpdatePosition(); and finally my fixed time step is declared in the constructor of my Game1.cs Class: /// <summary> /// The main game constructor. /// </summary> public Game1() { IsFixedTimeStep = slow = false; } So my question is: Where do I place the MoveTimer or elapsedTime, so that my bat will slow down accordingly?

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  • No wireless connection using a conceptronic c54i (RT2561/RT61 rev B)

    - by jrosell
    Detected but not working. New install on ubuntu 11.10 using coneptronic C54Ri. As documentation says it uses Ralink drivers.... Any ideas why my wireless does not work? $ lspci -nn | grep -i 'ralink' 01:05.0 Network controller: Ralink corp. RT2561/RT61 rev B 802.11g ifconfig eth0 Link encap:Ethernet HWaddr 00:1e:90:e5:af:13 inet addr:192.168.0.197 Bcast:192.168.0.255 Mask:255.255.255.0 inet6 addr: fe80::21e:90ff:fee5:af13/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:28361 errors:0 dropped:0 overruns:0 frame:0 TX packets:16858 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:39812172 (39.8 MB) TX bytes:1633405 (1.6 MB) Interrupt:43 Base address:0xc000 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host UP LOOPBACK RUNNING MTU:16436 Metric:1 RX packets:80 errors:0 dropped:0 overruns:0 frame:0 TX packets:80 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:0 RX bytes:6608 (6.6 KB) TX bytes:6608 (6.6 KB) iwconfig wlan0 wlan0 IEEE 802.11abg ESSIDff/any Mode:Managed Access Point: Not-Associated Tx-Power=0 dBm Retry long limit:7 RTS thrff Fragment thrff Power Managementff lsmod | grep rt rt61pci 27493 0 crc_itu_t 12627 1 rt61pci rt2x00pci 14202 1 rt61pci rt2x00lib 48114 2 rt61pci,rt2x00pci mac80211 272785 2 rt2x00pci,rt2x00lib cfg80211 172392 2 rt2x00lib,mac80211 eeprom_93cx6 12653 1 rt61pci parport_pc 32114 1 parport 40930 3 ppdev,parport_pc,lp lsmod | grep rt [ 2497.816989] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.827112] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.837430] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.847528] phy0 -> rt2x00pci_regbusy_read: Error - Indirect register access failed: offset=0x0000308c, value=0xffffffff [ 2497.847632] phy0 -> rt61pci_wait_bbp_ready: Error - BBP register access faile d, aborting. [ 2497.847637] phy0 -> rt61pci_set_device_state: Error - Device failed to enter state 4 (-5). sudo lshw -C network *-network DISABLED description: Wireless interface product: RT2561/RT61 rev B 802.11g vendor: Ralink corp. physical id: 5 bus info: pci@0000:01:05.0 logical name: wlan0 version: 00 serial: fa:b8:14:58:62:35 width: 32 bits clock: 33MHz capabilities: pm cap_list ethernet physical wireless configuration: broadcast=yes driver=rt61pci driverversion=3.0.0-12-generic firmware=0.8 latency=0 link=no multicast=yes wireless=IEEE 802.11abg resources: irq:16 memory:fdef8000-fdefffff iwlist scan lo Interface doesn't support scanning. eth0 Interface doesn't support scanning. wlan0 Failed to read scan data : Network is down uname -mr 3.0.0-12-generic i686 Edit 1 $ rfkill list all 0: phy0: Wireless LAN Soft blocked: no Hard blocked: no On reboot, sudo lshw -C network returns network is ok. Hovever, WPA keeps on asking the wireless key

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  • How to show pending messages using WLST?

    - by lmestre
    Here are the steps: 1. . ./setDomainEnv.sh2. java weblogic.WLST3. connect('weblogic','welcome1','t3://localhost:7001')4. domainRuntime()5. cd('ServerRuntimes/MS1/JMSRuntime/MS1.jms/JMSServers/JMSServer1/Destinations/JMSModule1!Queue1')6. cursor1=cmo.getMessages('true',9999999,10)                                                 **String(selector),Integer(timeout),Integer(state)7. msgs = cmo.getNext(cursor1, 10)                  ** This step gets 10 messages, you can call again cmo.getNext(cursor1, 10) to get the next 10 msgs8. print(msgs)My assumption, is that you had created:a. Managed Server MS1.b. JMS Server JMSServer1.c. Module called JMSModule1.d. Inside of JMSModule1, a Queue called Queue1.If you read my previous post:How to get Messages Pending Count from a Queue using WLST? https://blogs.oracle.com/LuzMestre/entry/how_to_get_messages_pendingYou can see that both are very similar.  Sometimes it is difficult to get a WLST Script sample, but you can use ls() function to know about other functionalities you don't have a sample code.***Until step 5, nothing new comparing to my previous post.5. cd('ServerRuntimes/MS1/JMSRuntime/MS1.jms/JMSServers/JMSServer1/Destinations/JMSModule1!Queue1')6. ls()You will see, MessagesPendingCount, getMessages along a lot of other functionalities available in this Queue. e.g, you can see:-r-x   getMessages                                  String : String(selector),Integer(timeout),Integer(state)Here you can check the complete MBean Reference:http://docs.oracle.com/cd/E23943_01/apirefs.1111/e13951/core/index.htmlSee JMSDestinationRuntimeMBean.Enjoy!

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  • Using Query Classes With NHibernate

    - by Liam McLennan
    Even when using an ORM, such as NHibernate, the developer still has to decide how to perform queries. The simplest strategy is to get access to an ISession and directly perform a query whenever you need data. The problem is that doing so spreads query logic throughout the entire application – a clear violation of the Single Responsibility Principle. A more advanced strategy is to use Eric Evan’s Repository pattern, thus isolating all query logic within the repository classes. I prefer to use Query Classes. Every query needed by the application is represented by a query class, aka a specification. To perform a query I: Instantiate a new instance of the required query class, providing any data that it needs Pass the instantiated query class to an extension method on NHibernate’s ISession type. To query my database for all people over the age of sixteen looks like this: [Test] public void QueryBySpecification() { var canDriveSpecification = new PeopleOverAgeSpecification(16); var allPeopleOfDrivingAge = session.QueryBySpecification(canDriveSpecification); } To be able to query for people over a certain age I had to create a suitable query class: public class PeopleOverAgeSpecification : Specification<Person> { private readonly int age; public PeopleOverAgeSpecification(int age) { this.age = age; } public override IQueryable<Person> Reduce(IQueryable<Person> collection) { return collection.Where(person => person.Age > age); } public override IQueryable<Person> Sort(IQueryable<Person> collection) { return collection.OrderBy(person => person.Name); } } Finally, the extension method to add QueryBySpecification to ISession: public static class SessionExtensions { public static IEnumerable<T> QueryBySpecification<T>(this ISession session, Specification<T> specification) { return specification.Fetch( specification.Sort( specification.Reduce(session.Query<T>()) ) ); } } The inspiration for this style of data access came from Ayende’s post Do You Need a Framework?. I am sick of working through multiple layers of abstraction that don’t do anything. Have you ever seen code that required a service layer to call a method on a repository, that delegated to a common repository base class that wrapped and ORMs unit of work? I can achieve the same thing with NHibernate’s ISession and a single extension method. If you’re interested you can get the full Query Classes example source from Github.

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