Search Results

Search found 40757 results on 1631 pages for 'inferred type'.

Page 383/1631 | < Previous Page | 379 380 381 382 383 384 385 386 387 388 389 390  | Next Page >

  • Unit Testing Framework for XQuery

    - by Knut Vatsendvik
    This posting provides a unit testing framework for XQuery using Oracle Service Bus. It allows you to write a test case to run your XQuery transformations in an automated fashion. When the test case is run, the framework returns any differences found in the response. The complete code sample with install instructions can be downloaded from here. Writing a Unit Test You start a new Test Case by creating a Proxy Service from Workshop that comes with Oracle Service Bus. In the General Configuration page select Service Type to be Messaging Service           In the Message Type Configuration page link both the Request & Response Message Type to the TestCase element of the UnitTest.xsd schema                 The TestCase element consists of the following child elements The ID and optional Name element is simply used for reference. The Transformation element is the XQuery resource to be executed. The Input elements represents the input to run the XQuery with. The Output element represents the expected output. These XML documents are “also” represented as an XQuery resource where the XQuery function takes no arguments and returns the XML document. Why not pass the test data with the TestCase? Passing an XML structure in another XML structure is not very easy or at least not very human readable. Therefore it was chosen to represent the test data as an loadable resource in the OSB. However you are free to go ahead with another approach on this if wanted. The XMLDiff elements represents any differences found. A sample on input is shown here. Modeling the Message Flow Then the next step is to model the message flow of the Proxy Service. In the Request Pipeline create a stage node that loads the test case input data.      For this, specify a dynamic XQuery expression that evaluates at runtime to the name of a pre-registered XQuery resource. The expression is of course set by the input data from the test case.           Add a Run stage node. Assign the result of the XQuery, that is to be run, to a context variable. Define a mapping for each of the input variables added in previous stage.     Add a Compare stage. Like with the input data, load the expected output data. Do a compare using XMLDiff XQuery provided where the first argument is the loaded output test data, and the second argument the result from the Run stage. Any differences found is replaced back into the test case XMLDiff element. In case of any unexpected failure while processing, add an Error Handler to the Pipeline to capture the fault. To pass back the result add the following Insert action In the Response Pipeline. A sample on output is shown here.

    Read the article

  • No sound lenovo t60 alsa ad1981 iec958

    - by Nate
    Any help on getting the sound to come through my lenovo t60 build in speakers, headphones, or mic would greatly be appreciated. The three buttons to increase, decrease sound seem to work. Bios has sound card enabled and the buttons beep when pressed. When going to Utube or playing music, no sound is heard. Thanks Nate Feb 23 - Didn't see anything specific in the sys logs with Rhythmbox when connecting my ipod. Rhythmbox is playing, but still no sound. Here is the syslog details for today. Output is set to analog output. Feb 23 17:42:32 itgis01398 rsyslogd: [origin software="rsyslogd" swVersion="4.2.0" x-pid="824" x-info="http://www.rsyslog.com"] rsyslogd was HUPed, type 'lightweight'. Feb 23 17:42:33 itgis01398 rsyslogd: [origin software="rsyslogd" swVersion="4.2.0" x-pid="824" x-info="http://www.rsyslog.com"] rsyslogd was HUPed, type 'lightweight'. Feb 23 17:42:49 itgis01398 anacron[968]: Job `cron.daily' terminated Feb 23 17:42:49 itgis01398 anacron[968]: Job `cron.weekly' started Feb 23 17:42:49 itgis01398 anacron[12067]: Updated timestamp for job `cron.weekly' to 2011-02-23 Feb 23 17:42:53 itgis01398 anacron[968]: Job `cron.weekly' terminated Feb 23 17:42:53 itgis01398 anacron[968]: Normal exit (2 jobs run) Feb 23 18:01:19 itgis01398 kernel: [ 2731.324067] usb 1-5: new high speed USB device using ehci_hcd and address 3 Feb 23 18:01:19 itgis01398 kernel: [ 2731.482879] Initializing USB Mass Storage driver... Feb 23 18:01:19 itgis01398 kernel: [ 2731.483061] usb-storage 1-5:1.0: Quirks match for vid 05ac pid 1205: 10 Feb 23 18:01:19 itgis01398 kernel: [ 2731.483116] scsi6 : usb-storage 1-5:1.0 Feb 23 18:01:19 itgis01398 kernel: [ 2731.483306] usbcore: registered new interface driver usb-storage Feb 23 18:01:19 itgis01398 kernel: [ 2731.483310] USB Mass Storage support registered. Feb 23 18:01:20 itgis01398 kernel: [ 2732.481116] scsi 6:0:0:0: Direct-Access Apple iPod 1.62 PQ: 0 ANSI: 0 Feb 23 18:01:20 itgis01398 kernel: [ 2732.482466] sd 6:0:0:0: Attached scsi generic sg2 type 0 Feb 23 18:01:20 itgis01398 kernel: [ 2732.485095] sd 6:0:0:0: [sdb] Adjusting the sector count from its reported value: 7999488 Feb 23 18:01:20 itgis01398 kernel: [ 2732.485110] sd 6:0:0:0: [sdb] 7999487 512-byte logical blocks: (4.09 GB/3.81 GiB) Feb 23 18:01:20 itgis01398 kernel: [ 2732.487933] sd 6:0:0:0: [sdb] Write Protect is off Feb 23 18:01:20 itgis01398 kernel: [ 2732.487941] sd 6:0:0:0: [sdb] Mode Sense: 64 00 00 08 Feb 23 18:01:20 itgis01398 kernel: [ 2732.487947] sd 6:0:0:0: [sdb] Assuming drive cache: write through Feb 23 18:01:20 itgis01398 kernel: [ 2732.489927] sd 6:0:0:0: [sdb] Adjusting the sector count from its reported value: 7999488 Feb 23 18:01:20 itgis01398 kernel: [ 2732.491150] sd 6:0:0:0: [sdb] Assuming drive cache: write through Feb 23 18:01:20 itgis01398 kernel: [ 2732.491163] sdb: sdb1 sdb2 Feb 23 18:01:20 itgis01398 kernel: [ 2732.510428] sd 6:0:0:0: [sdb] Adjusting the sector count from its reported value: 7999488 Feb 23 18:01:20 itgis01398 kernel: [ 2732.511288] sd 6:0:0:0: [sdb] Assuming drive cache: write through Feb 23 18:01:20 itgis01398 kernel: [ 2732.511297] sd 6:0:0:0: [sdb] Attached SCSI removable disk Feb 23 18:01:21 itgis01398 kernel: [ 2733.746675] FAT: invalid media value (0x2f) Feb 23 18:01:21 itgis01398 kernel: [ 2733.746682] VFS: Can't find a valid FAT filesystem on dev sdb1. Feb 23 18:01:22 itgis01398 upstart-udev-bridge[330]: Env must be KEY=VALUE pairs Feb 23 18:02:07 itgis01398 kernel: [ 2780.115826] sd 6:0:0:0: [sdb] Unhandled sense code Feb 23 18:02:07 itgis01398 kernel: [ 2780.115835] sd 6:0:0:0: [sdb] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Feb 23 18:02:07 itgis01398 kernel: [ 2780.115844] sd 6:0:0:0: [sdb] Sense Key : Medium Error [current] Feb 23 18:02:07 itgis01398 kernel: [ 2780.115855] Info fld=0x0 Feb 23 18:02:07 itgis01398 kernel: [ 2780.115859] sd 6:0:0:0: [sdb] Add. Sense: Unrecovered read error Feb 23 18:02:07 itgis01398 kernel: [ 2780.115870] sd 6:0:0:0: [sdb] CDB: Read(10): 28 00 00 08 fd e9 00 00 f0 00 Feb 23 18:02:07 itgis01398 kernel: [ 2780.115892] end_request: I/O error, dev sdb, sector 589289 Feb 23 18:02:49 itgis01398 kernel: [ 2821.351464] sd 6:0:0:0: [sdb] Unhandled sense code Feb 23 18:02:49 itgis01398 kernel: [ 2821.351473] sd 6:0:0:0: [sdb] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Feb 23 18:02:49 itgis01398 kernel: [ 2821.351482] sd 6:0:0:0: [sdb] Sense Key : Medium Error [current] Feb 23 18:02:49 itgis01398 kernel: [ 2821.351493] Info fld=0x0 Feb 23 18:02:49 itgis01398 kernel: [ 2821.351497] sd 6:0:0:0: [sdb] Add. Sense: No additional sense information Feb 23 18:02:49 itgis01398 kernel: [ 2821.351507] sd 6:0:0:0: [sdb] CDB: Read(10): 28 00 00 08 fe d9 00 00 10 00 Feb 23 18:02:49 itgis01398 kernel: [ 2821.351530] end_request: I/O error, dev sdb, sector 589529 Feb 23 18:17:01 itgis01398 CRON[12709]: (root) CMD ( cd / && run-parts --report /etc/cron.hourly) volume is all of the way up.

    Read the article

  • Get Exchange Online Mailbox Size in GB

    - by Brian Jackett
    As mentioned in my previous post I was recently working with a customer to get started with Exchange Online PowerShell commandlets.  In this post I wanted to follow up and show one example of a difference in output from commandlets in Exchange 2010 on-premises vs. Exchange Online.   Problem    The customer was interested in getting the size of mailboxes in GB.  For Exchange on-premises this is fairly easy.  A fellow PFE Gary Siepser wrote an article explaining how to accomplish this (click here).  Note that Gary’s script will not work when remoting from a local machine that doesn’t have the Exchange object model installed.  A similar type of scenario exists if you are executing PowerShell against Exchange Online.  The data type for TotalItemSize  being returned (ByteQuantifiedSize) exists in the Exchange namespace.  If the PowerShell session doesn’t have access to that namespace (or hasn’t loaded it) PowerShell works with an approximation of that data type.    The customer found a sample script on this TechNet article that they attempted to use (minor edits by me to fit on page and remove references to deleted item size.)   Get-Mailbox -ResultSize Unlimited | Get-MailboxStatistics | Select DisplayName,StorageLimitStatus, ` @{name="TotalItemSize (MB)"; expression={[math]::Round( ` ($_.TotalItemSize.Split("(")[1].Split(" ")[0].Replace(",","")/1MB),2)}}, ` ItemCount | Sort "TotalItemSize (MB)" -Descending | Export-CSV "C:\My Documents\All Mailboxes.csv" -NoTypeInformation     The script is targeted to Exchange 2010 but fails for Exchange Online.  In Exchange Online when referencing the TotalItemSize property though it does not have a Split method which ultimately causes the script to fail.   Solution    A simple solution would be to add a call to the ToString method off of the TotalItemSize property (in bold on line 5 below).   Get-Mailbox -ResultSize Unlimited | Get-MailboxStatistics | Select DisplayName,StorageLimitStatus, ` @{name="TotalItemSize (MB)"; expression={[math]::Round( ` ($_.TotalItemSize.ToString().Split("(")[1].Split(" ")[0].Replace(",","")/1MB),2)}}, ` ItemCount | Sort "TotalItemSize (MB)" -Descending | Export-CSV "C:\My Documents\All Mailboxes.csv" -NoTypeInformation      This fixes the script to run but the numerous string replacements and splits are an eye sore to me.  I attempted to simplify the string manipulation with a regular expression (more info on regular expressions in PowerShell click here).  The result is a workable script that does one nice feature of adding a new member to the mailbox statistics called TotalItemSizeInBytes.  With this member you can then convert into any byte level (KB, MB, GB, etc.) that suits your needs.  You can download the full version of this script below (includes commands to connect to Exchange Online session). $UserMailboxStats = Get-Mailbox -RecipientTypeDetails UserMailbox ` -ResultSize Unlimited | Get-MailboxStatistics $UserMailboxStats | Add-Member -MemberType ScriptProperty -Name TotalItemSizeInBytes ` -Value {$this.TotalItemSize -replace "(.*\()|,| [a-z]*\)", ""} $UserMailboxStats | Select-Object DisplayName,@{Name="TotalItemSize (GB)"; ` Expression={[math]::Round($_.TotalItemSizeInBytes/1GB,2)}}   Conclusion    Moving from on-premises to the cloud with PowerShell (and PowerShell remoting in general) can sometimes present some new challenges due to what you have access to.  This means that you must always test your code / scripts.  I still believe that not having to physically RDP to a server is a huge gain over some of the small hurdles you may encounter during the transition.  Scripting is the future of administration and makes you more valuable.  Hopefully this script and the concepts presented help you be a better admin / developer.         -Frog Out     Links The Get-MailboxStatistics Cmdlet, the TotalitemSize Property, and that pesky little “b” http://blogs.technet.com/b/gary/archive/2010/02/20/the-get-mailboxstatistics-cmdlet-the-totalitemsize-property-and-that-pesky-little-b.aspx   View Mailbox Sizes and Mailbox Quotas Using Windows PowerShell http://technet.microsoft.com/en-us/exchangelabshelp/gg576861#ViewAllMailboxes   Regular Expressions with Windows PowerShell http://www.regular-expressions.info/powershell.html   “I don’t always test my code…” image http://blogs.pinkelephant.com/images/uploads/conferences/I-dont-always-test-my-code-But-when-I-do-I-do-it-in-production.jpg   The One Thing: Brian Jackett and SharePoint 2010 http://www.youtube.com/watch?v=Sg_h66HMP9o

    Read the article

  • SQL SERVER – Select the Most Optimal Backup Methods for Server

    - by pinaldave
    Backup and Restore are very interesting concepts and one should be very much with the concept if you are dealing with production database. One never knows when a natural disaster or user error will surface and the first thing everybody wants is to get back on point in time when things were all fine. Well, in this article I have attempted to answer a few of the common questions related to Backup methodology. How to Select a SQL Server Backup Type In order to select a proper SQL Server backup type, a SQL Server administrator needs to understand the difference between the major backup types clearly. Since a picture is worth a thousand words, let me offer it to you below. Select a Recovery Model First The very first question that you should ask yourself is: Can I afford to lose at least a little (15 min, 1 hour, 1 day) worth of data? Resist the temptation to save it all as it comes with the overhead – majority of businesses outside finances can actually afford to lose a bit of data. If your answer is YES, I can afford to lose some data – select a SIMPLE (default) recovery model in the properties of your database, otherwise you need to select a FULL recovery model. The additional advantage of the Full recovery model is that it allows you to restore the data to a specific point in time vs to only last backup time in the Simple recovery model, but it exceeds the scope of this article Backups in SIMPLE Recovery Model In SIMPLE recovery model you can select to do just Full backups or Full + Differential. Full Backup This is the simplest type of backup that contains all information needed to restore the database and should be your first choice. It is often sufficient for small databases, but note that it makes a big impact on the performance of your database Full + Differential Backup After Full, Differential backup picks up all of the changes since the last Full backup. This means if you made Full, Diff, Diff backup – the last Diff backup contains all of the changes and you don’t need the previous Differential backup. Differential backup is obviously smaller and carries less performance overhead Backups in FULL Recovery Model In FULL recovery model you can select Full + Transaction Log or Full + Differential + Transaction Log backup. You have to create Transaction Log backup, because at that time the log is being truncated. Otherwise your Transaction Log will grow uncontrollably. Full + Transaction Log Backup You would always need to perform a Full backup first. Then a series of Transaction log backup. Note that (in contrast to Differential) you need ALL transactions to log since the last Full of Diff backup to properly restore. Transaction log backups have the smallest performance overhead and can be performed often. Full + Differential + Transaction Log Backup If you want to ease the performance overhead on your server, you can replace some of the Full backup in the previous scenario with Differential. You restore scenario would start from Full, then the Last Differential, then all of the remaining transactions log backups Typical backup Scenarios You may say “Well, it is all nice – give me the examples now”. As you may already know, my favorite SQL backup software is SQLBackupAndFTP. If you go to Advanced Backup Schedule form in this program and click “Load a typical backup plan…” link, it will give you these scenarios that I think are quite common – see the image below. The Simplest Way to Schedule SQL Backups I hate to repeat myself, but backup scheduling in SQL agent leaves a lot to be desired. I do not know the simple way to schedule your SQL server backups than in SQLBackupAndFTP – see the image below. The whole backup scheduling with compression, encryption and upload to a Network Folder / HDD / NAS Drive / FTP / Dropbox / Google Drive / Amazon S3 takes just a few minutes – see my previous post for the review. Final Words This post offered an explanation for major backup types only. For more complicated scenarios or to research other options as usually go to MSDN. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: PostADay, SQL, SQL Authority, SQL Backup and Restore, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

    Read the article

  • Is there a good [and modern] reason to not have static HTML pages with AJAX content , rather than generate pages?

    - by user1725
    Assumptions: We don't care about IE6, and Noscript users. Lets pretend we have the following design concept: All your pages are HTML/CSS that create the ascetics, layout, colours, general design related things. Lets pretend this basic code below is that: <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Strict//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-strict.dtd"> <html> <head> <link href="/example.css" rel="stylesheet" type="text/css"/> <script src="example.js" type="text/javascript"></script> <head> <body> <div class="left"> </div> <div class="mid"> </div> <div class="right"> </div> </body> </html> Which in theory should produce, with the right CSS, three vertical columns on the web page. Now, here's the root of the question, what are the serious advantages and/or disadvantages of loading the content of these columns (lets assume they are all indeed dynamic content, not static) via AJAX requests, or have the content pre-set with a scripting language? So for instance, we would have, in the AJAX example, lets asume jquery is used on-load: //Multiple http requests $("body > div.left").load("./script.php?content=news"); $("body > div.right").load("./script.php?content=blogs"); $("body > div.mid").load("./script.php?content=links"); OR--- //Single http request $.ajax({ url: './script.php?content=news|blogs|links', method: 'json', type: 'text', success: function (data) { $("body > div.left").html(data.news); $("body > div.right").html(data.blogs); $("body > div.mid").html(data.links); } }) Verses doing this: <body> <div class="left"> <?php echo function_returning_news(); ?> </div> <div class="mid"> <?php echo function_returning_blogs(); ?> </div> <div class="right"> <?php echo function_returning_links(); ?> </div> </body> I'm personally thinking right now that doing static HTML pages is a better method, my reasoning is: I've separated my data, logic, and presentation (ie, "MVC") code. I can make changes to one without others. Browser caches mean I'm just getting server load mostly for the content, not the presentation wrapped around it. I could turn my "script.php" into a more robust API for the website. But I'm not certain or clear that these are legitimately good reasons, and I'm not confidently aware of other issues that could happen, so I would like to know the pros-and-cons, so to speak.

    Read the article

  • How to Add Your Gmail Account to Outlook 2013 Using IMAP

    - by Lori Kaufman
    If you use Outlook to check and manage your email, you can easily use it to check your Gmail account as well. You can setup your Gmail account to allow you to synchronize email across multiple machines using email clients instead of a browser. We will show you how to use IMAP in your Gmail account so you can synchronize your Gmail account across multiple machines, and then how to add your Gmail account to Outlook 2013. To setup your Gmail account to use IMAP, sign in to your Gmail account and go to Mail. Click the Settings button in the upper, right corner of the window and select Settings from the drop-down menu. On the Settings screen, click Forwarding and POP/IMAP. Scroll down to the IMAP Access section and select Enable IMAP. Click Save Changes at the bottom of the screen. Close your browser and open Outlook. To begin adding your Gmail account, click the File tab. On the Account Information screen, click Add Account. On the Add Account dialog box, you can choose the E-mail Account option which automatically sets up your Gmail account in Outlook. To do this enter your name, email address, and the password for your Gmail account twice. Click Next. The progress of the setup displays. The automatic process may or may not work. If the automatic process fails, select Manual setup or additional server types, instead of E-mail Account, and click Next. On the Choose Service screen, select POP or IMAP and click Next. On the POP and IMAP Account Settings enter the User, Server, and Logon Information. For the Server Information, select IMAP from the Account Type drop-down list and enter the following for the incoming and outgoing server information: Incoming mail server: imap.googlemail.com Outgoing mail server (SMTP): smtp.googlemail.com Make sure you enter your full email address for the User Name and select Remember password if you want Outlook to automatically log you in when checking email. Click More Settings. On the Internet E-mail Settings dialog box, click the Outgoing Server tab. Select the My outgoing server (SMTP) requires authentication and make sure the Use same settings as my incoming mail server option is selected. While still in the Internet E-mail Settings dialog box, click the Advanced tab. Enter the following information: Incoming server: 993 Incoming server encrypted connection: SSL Outgoing server encrypted connection TLS Outgoing server: 587 NOTE: You need to select the type of encrypted connection for the outgoing server before entering 587 for the Outgoing server (SMTP) port number. If you enter the port number first, the port number will revert back to port 25 when you change the type of encrypted connection. Click OK to accept your changes and close the Internet E-mail Settings dialog box. Click Next. Outlook tests the accounts settings by logging into the incoming mail server and sending a test email message. When the test is finished, click Close. You should see a screen saying “You’re all set!”. Click Finish. Your Gmail address displays in the account list on the left with any other email addresses you have added to Outlook. Click the Inbox to see what’s in your Inbox in your Gmail account. Because you’re using IMAP in your Gmail account and you used IMAP to add the account to Outlook, the messages and folders in Outlook reflect what’s in your Gmail account. Any changes you make to folders and any time you move email messages among folders in Outlook, the same changes are made in your Gmail account, as you will see when you log into your Gmail account in a browser. This works the other way as well. Any changes you make to the structure of your account (folders, etc.) in a browser will be reflected the next time you log into your Gmail account in Outlook.     

    Read the article

  • ResourceSerializable: an alternate to ORM and ActiveRecord

    - by Levi Morrison
    A few opinionated reasons I don't like the traditional ORM and ActiveRecord patterns: They work only with a database. Sometimes I'm dealing with objects from an API and other objects from a database. All the implementations I have seen don't allow for that. Feel free to clue me in if I'm wrong on this. They are brittle. Changes in the database will likely break your implemenation. Some implementations can help reduce this, but a few of the ones I've seen don't. Their very design is influenced by the database. If I want to switch to using an API, I'll have to redesign the object to get it to work (likely). It seems to violate the single-responsibility pattern. They know what they are and how they act, but they also know how they are created, destroyed and saved? Seems a bit much. What about an approach that is somewhat more familiar in PHP: implementing an interface? In php 5.4, we'll have the JsonSerializable interface that defines the data to be json_encoded, so users will become accustomed to this type of thing. What if there was a ResourceSerializable interface? This is still an ORM by name, but certainly not by tradition. interface ResourceSerializable { /** * Returns the id that identifies the resource. */ function resourceId(); /** * Returns the 'type' of the resource. */ function resourceType(); /** * Returns the data to be serialized. */ function resourceSerialize(); } Things might be poorly named, I'll take suggestions. Notes: ResourceId will work for API's and databases. As long as your primary key in the database is the same as the resource ID in the API, there is no conflict. All of the API's I've worked with have a unique ID for the resource, so I don't see any issues there. ResourceType is the group or type associated with the resource. You can use this to map the resource to an API call or a database table. If the ResourceType was person, it could map to /api/1/person/{resourceId} and the table persons (or people, if it's smart enough). resourceSerialize() returns the data to be stored. Keys would identify API parameters and database table columns. This also seems easier to test than ActiveRecord / Orm implemenations. I haven't done much automated testing on traditional ActiveRecord/ORM implemenations, so this is merely a guess. But it seems that I being able to create objects independently of the library helps me. I don't have to use load() to get an existing resource, I can simply create one and set all the right properties. This is not so easy in the ActiveRecord / Orm implemenations I've dealt with. Downsides: You need another object to serialize it. This also means you have more code in general as you have to use more objects. You have to map resource types to API calls and database tables. This is even more work, but some ORMs and ActiveRecord implementations require you to map objects to table names anyway. Are there other downsides that you see? Does this seem feasible to you? How would you improve it? Note: I almost asked this on StackOverflow because it might be too vague for their standards, but I'm still not really familiar with programmers.stackexchange.com, so please help me improve my question if it doesn't shape up to standards here.

    Read the article

  • Surviving MATLAB and R as a Hardcore Programmer

    - by dsimcha
    I love programming in languages that seem geared towards hardcore programmers. (My favorites are Python and D.) MATLAB is geared towards engineers and R is geared towards statisticians, and it seems like these languages were designed by people who aren't hardcore programmers and don't think like hardcore programmers. I always find them somewhat awkward to use, and to some extent I can't put my finger on why. Here are some issues I have managed to identify: (Both): The extreme emphasis on vectors and matrices to the extent that there are no true primitives. (Both): The difficulty of basic string manipulation. (Both): Lack of or awkwardness in support for basic data structures like hash tables and "real", i.e. type-parametric and nestable, arrays. (Both): They're really, really slow even by interpreted language standards, unless you bend over backwards to vectorize your code. (Both): They seem to not be designed to interact with the outside world. For example, both are fairly bulky programs that take a while to launch and seem to not be designed to make simple text filter programs easy to write. Furthermore, the lack of good string processing makes file I/O in anything but very standard forms near impossible. (Both): Object orientation seems to have a very bolted-on feel. Yes, you can do it, but it doesn't feel much more idiomatic than OO in C. (Both): No obvious, simple way to get a reference type. No pointers or class references. For example, I have no idea how you roll your own linked list in either of these languages. (MATLAB): You can't put multiple top level functions in a single file, encouraging very long functions and cut-and-paste coding. (MATLAB): Integers apparently don't exist as a first class type. (R): The basic builtin data structures seem way too high level and poorly documented, and never seem to do quite what I expect given my experience with similar but lower level data structures. (R): The documentation is spread all over the place and virtually impossible to browse or search. Even D, which is often knocked for bad documentation and is still fairly alpha-ish, is substantially better as far as I can tell. (R): At least as far as I'm aware, there's no good IDE for it. Again, even D, a fairly alpha-ish language with a small community, does better. In general, I also feel like MATLAB and R could be easily replaced by plain old libraries in more general-purpose langauges, if sufficiently comprehensive libraries existed. This is especially true in newer general purpose languages that include lots of features for library writers. Why do R and MATLAB seem so weird to me? Are there any other major issues that you've noticed that may make these languages come off as strange to hardcore programmers? When their use is necessary, what are some good survival tips? Edit: I'm seeing one issue from some of the answers I've gotten. I have a strong personal preference, when I analyze data, to have one script that incorporates the whole pipeline. This implies that a general purpose language needs to be used. I hate having to write a script to "clean up" the data and spit it out, then another to read it back in a completely different environment, etc. I find the friction of using MATLAB/R for some of my work and a completely different language with a completely different address space and way of thinking for the rest to be a huge source of friction. Furthermore, I know there are glue layers that exist, but they always seem to be horribly complicated and a source of friction.

    Read the article

  • Windows 8 Camp&ndash;Ways to Prepare

    - by Lori Lalonde
    When Windows 8 was announced at the BUILD conference back in September, it created quite a buzz among the developer community. By the spring of 2012,  Windows 8 Developer Camps started popping up everywhere imaginable. I received a lot of questions from CTTDNUG members about whether or not we would be hosting one locally. If you recall my post about the Windows Phone/Azure Developer Workshop that CTTDNUG hosted back in March, you’ll remember that the biggest hurdle to overcome when planning this type of event was finding the right venue. It took some time, but I finally found a venue that was available and provided the prerequisites needed to ensure this camp is a success. I am very excited that CTTDNUG will be hosting a Windows 8 Camp this summer in the Kitchener/Waterloo area. In fact, it’s coming up in less than 2 weeks. Clearly other developers are excited as well, because our registration numbers show that the event is already 70% full! On top of that, I was fortunate enough to also book two well-known evangelists to present and teach at this full day developer camp: Andrei Marukovich and Atley Hunter. This was the icing on the cake. With the content provided by Microsoft, and two local experts that live and breathe Windows 8 development, I know that I, along with other developers that attend this event, will have the opportunity to maximize our learning potential and hit the ground running. If you plan on attending a Windows 8 Developer Camp soon, and want to ensure you get the most “bang for your buck” (figuratively speaking, since these camps are free), there are some things you can do to prepare before the big day: 1) Install the prerequisites on your own device before the big day I can’t stress this enough. Otherwise, you will be spending valuable time during the hands-on period downloading and installing what is needed, rather than digging into the development and using that time to ask the experts on-hand about programming challenges, issues, questions you may have with respect to your development. Prerequisites: Windows 8 Release Preview Visual Studio 2012 RC Download the Windows 8 SDK Samples 2) Purchase, download, and read Charles Petzold’s newest book:  Programming Windows 6th Edition This is a great introduction to the type of content you will be learning about during the camp. Doing some light reading beforehand might raise some questions about the concepts discussed in the book, which will give you the opportunity to write them down and bring them with you to the camp. The experts on hand will be able to answer them for you. 3) Make use of the freebies that are available Telerik has recently released a preview of their RadControls for Metro. You can sign up to receive a license code to give you access to install the preview for free and start playing around with it. Syncfusion also offers a free download of their Metro Studio package, which is a collection of metro style icons that you can customize and use in your own applications. Last but not least, once you’ve installed the Windows 8 Release Preview on your own device, go to the Windows 8 Store and download a handful of the free apps that are available. Testing out other Metro apps may give you ideas of what you can do in your own apps and analyze what features you like: application flow, type of animations used, concepts that were leveraged, how live tiles were used, etc. I hope you found these tips to be useful as you embark on a new development journey! Although this post focused on how to prepare for a Windows 8 camp, the same ideas are there whichever developer camp/workshop/event you attend. Learning does not begin and end on the day of the event. Attending a developer camp is just one step of many to master whatever technology you are interested in. It is a continuous process, which is fully maximized when you do your homework beforehand, actively participate during,  and follow up by putting what you learned to practice afterwards. Happy coding!

    Read the article

  • Problem with ebay AddItem API call [migrated]

    - by user1323572
    I am totally new to any sort of API application. Right now I am creating a listing application to list items on E-bay India site. API version being used is 767, sandbox url is https://api.sandbox.ebay.com/wsapi. I have sandbox account for ebay(buyer/seller) and developer account. I am getting error saying: 1) Sales Tax / VAT was dropped from the listing as per new sales tax / VAT policy. The items will be listed successfully, you may revise the listing to specify all inclusive price. 2) You have either not registered or are having problem with your payment method registration. ItemType type = new ItemType(); type.PaymentMethods = new BuyerPaymentMethodCodeTypeCollection(); type.PaymentMethods.Add(BuyerPaymentMethodCodeType.PaisaPayAccepted); Also do I have to specify taxation for each state? For VAT and shipping details here's my snippet: private ShippingDetailsType getShippingDetails() { // Shipping details. ShippingDetailsType sd = new ShippingDetailsType(); SalesTaxType salesTax = new SalesTaxType(); ReadSettings rs = new ReadSettings(); rs.GetSettings(); salesTax.SalesTaxPercent = 12f; salesTax.SalesTaxState = "MH"; SalesTaxType s = new SalesTaxType(); salesTax.ShippingIncludedInTax = true; salesTax.ShippingIncludedInTaxSpecified = true; sd.ApplyShippingDiscount = true; AmountType at = new AmountType(); at.Value = 2.8; at.currencyID = CurrencyCodeType.INR; sd.InsuranceFee = at; sd.InsuranceOption = InsuranceOptionCodeType.NotOffered; sd.PaymentInstructions = "These are my instructions."; VATDetailsType vd = new VATDetailsType(); vd.BusinessSeller = false; vd.BusinessSellerSpecified = false; vd.RestrictedToBusiness = false; vd.RestrictedToBusinessSpecified = false; vd.VATID = "VATSNO1234567890"; vd.VATPercent = 12f; vd.VATPercentSpecified = true; vd.VATSite = "None"; sd.ShippingType = ShippingTypeCodeType.Flat; // ShippingServiceOptionsType st1 = new ShippingServiceOptionsType(); sd.SalesTax = salesTax; st1.ShippingService = ShippingServiceCodeType.IN_Express.ToString(); at = new AmountType(); at.Value = 50; at.currencyID = CurrencyCodeType.INR; st1.ShippingServiceAdditionalCost = at; at = new AmountType(); at.Value = 50; at.currencyID = CurrencyCodeType.INR; st1.ShippingServiceCost = at; st1.ShippingServicePriority = 1; at = new AmountType(); at.Value = 1.0; at.currencyID = CurrencyCodeType.INR; st1.ShippingInsuranceCost = at; sd.ShippingServiceOptions = new ShippingServiceOptionsTypeCollection(new ShippingServiceOptionsType[] { st1 }); return sd; } Thank you for you efforts.

    Read the article

  • Can't connect to VPN on Ubuntu 12.04

    - by 12rad
    I'm having a lot of trouble connecting to VPN. This used to work on my machine, but i recently did an update and it's stopped working. I'm not sure what the problem is. My question is how do i debug this? I'm not able to narrow it down to a specific problem. This is what i get when i tail the syslogs. Would appreciate any help! Nov 6 23:42:52 meera NetworkManager[1137]: <info> Starting VPN service 'pptp'... Nov 6 23:42:52 meera NetworkManager[1137]: <info> VPN service 'pptp' started (org.freedesktop.NetworkManager.pptp), PID 6132 Nov 6 23:42:52 meera NetworkManager[1137]: <info> VPN service 'pptp' appeared; activating connections Nov 6 23:42:52 meera NetworkManager[1137]: <info> VPN plugin state changed: starting (3) Nov 6 23:42:52 meera NetworkManager[1137]: <info> VPN connection 'NAME VPN' (Connect) reply received. Nov 6 23:42:52 meera pppd[6136]: Plugin /usr/lib/pppd/2.4.5/nm-pptp-pppd-plugin.so loaded. Nov 6 23:42:52 meera pppd[6136]: pppd 2.4.5 started by root, uid 0 Nov 6 23:42:52 meera chat[6139]: timeout set to 15 seconds Nov 6 23:42:52 meera chat[6139]: abort on (NO CARRIER) Nov 6 23:42:52 meera chat[6139]: abort on (NO DIALTONE) Nov 6 23:42:52 meera chat[6139]: abort on (ERROR) Nov 6 23:42:52 meera chat[6139]: abort on (NO ANSWER) Nov 6 23:42:52 meera chat[6139]: abort on (BUSY) Nov 6 23:42:52 meera chat[6139]: abort on (Username/Password Incorrect) Nov 6 23:42:52 meera chat[6139]: send (AT^M) Nov 6 23:42:52 meera pptp[6138]: nm-pptp-service-6132 log[main:pptp.c:314]: The synchronous pptp option is NOT activated Nov 6 23:42:52 meera chat[6139]: expect (OK) Nov 6 23:42:52 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 1 'Start-Control-Connection-Request' Nov 6 23:42:53 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_disp:pptp_ctrl.c:739]: Received Start Control Connection Reply Nov 6 23:42:53 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_disp:pptp_ctrl.c:773]: Client connection established. Nov 6 23:42:53 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 7 'Outgoing-Call-Request' Nov 6 23:42:54 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_disp:pptp_ctrl.c:858]: Received Outgoing Call Reply. Nov 6 23:42:54 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_disp:pptp_ctrl.c:897]: Outgoing call established (call ID 0, peer's call ID 13077). Nov 6 23:42:54 meera pptp[6138]: nm-pptp-service-6132 warn[decaps_hdlc:pptp_gre.c:231]: The ppp mode is synchronous, yet no pptp --sync option is specified! Nov 6 23:43:07 meera chat[6139]: alarm Nov 6 23:43:07 meera chat[6139]: Failed Nov 6 23:43:07 meera pppd[6136]: Script chat -v -f /etc/ppp/chat-ztisp finished (pid 6139), status = 0x3 Nov 6 23:43:07 meera pppd[6136]: Connect script failed Nov 6 23:43:07 meera pppd[6136]: Waiting for 1 child processes... Nov 6 23:43:07 meera pppd[6136]: script /usr/sbin/pptp 204.197.218.90 --nolaunchpppd --loglevel 0 --logstring nm-pptp-service-6132, pid 6138 Nov 6 23:43:07 meera pptp[6138]: nm-pptp-service-6132 warn[decaps_hdlc:pptp_gre.c:204]: short read (-1): Input/output error Nov 6 23:43:07 meera pptp[6138]: nm-pptp-service-6132 warn[decaps_hdlc:pptp_gre.c:216]: pppd may have shutdown, see pppd log Nov 6 23:43:07 meera pptp[6143]: nm-pptp-service-6132 log[callmgr_main:pptp_callmgr.c:234]: Closing connection (unhandled) Nov 6 23:43:07 meera pppd[6136]: Script /usr/sbin/pptp 204.197.218.90 --nolaunchpppd --loglevel 0 --logstring nm-pptp-service-6132 finished (pid 6138), status = 0x0 Nov 6 23:43:07 meera pptp[6143]: nm-pptp-service-6132 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 12 'Call-Clear-Request' Nov 6 23:43:07 meera pptp[6143]: nm-pptp-service-6132 log[call_callback:pptp_callmgr.c:79]: Closing connection (call state) Nov 6 23:43:07 meera pppd[6136]: Exit. Nov 6 23:43:07 meera NetworkManager[1137]: <warn> VPN plugin failed: 1 Nov 6 23:43:07 meera NetworkManager[1137]: <info> VPN plugin state changed: stopped (6) Nov 6 23:43:07 meera NetworkManager[1137]: <info> VPN plugin state change reason: 0 Nov 6 23:43:07 meera NetworkManager[1137]: <warn> error disconnecting VPN: Could not process the request because no VPN connection was active.

    Read the article

  • OEG11gR2 integration with OES11gR2 Authorization with condition

    - by pgoutin
    Introduction This OES use-case has been defined originally by Subbu Devulapalli (http://accessmanagement.wordpress.com/).  Based on this OES museum use-case, I have developed the OEG11gR2 policy able to deal with the OES authorization with condition. From an OEG point of view, the way to deal with OES condition is to provide with the OES request some Environmental / Context Attributes.   Museum Use-Case  All painting in the museum have security sensors, an alarm goes off when a person comes too close a painting. The employee designated for maintenance needs to use their ID and disable the alarm before maintenance. You are the Security Administrator for the museum and you have been tasked with creating authorization policies to manage authorization for different paintings. Your first task is to understand how paintings are organized. Asking around, you are surprised to see that there isno formal process in place, so you need to start from scratch. the museum tracks the following attributes for each painting 1. Name of the work 2. Painter 3. Condition (good/poor) 4. Cost You compile the list of paintings  Name of Painting  Painter  Paint Condition  Cost  Mona Lisa  Leonardo da Vinci  Good  100  Magi  Leonardo da Vinci  Poor  40  Starry Night  Vincent Van Gogh  Poor  75  Still Life  Vincent Van Gogh  Good  25 Being a software geek who doesn’t (yet) understand art, you feel that price(or insurance price) of a painting is the most important criteria. So you feel that based on years-of-experience employees can be tasked with maintaining different paintings. You decide that paintings worth over 50 cost should be only handled by employees with over 20 years of experience and employees with less than 10 years of experience should not handle any painting. Lets us start with policy modeling. All paintings have a common set of attributes and actions, so it will be good to have them under a single Resource Type. Based on this resource type we will create the actual resources. So our high level model is: 1) Resource Type: Painting which has action manage and the following four attributes a) Name of the work b) Painter c) Condition (good/poor) d) Cost 2) To keep things simple lets use painting name for Resource name (in real world you will try to use some identifier which is unique, because in future we may end up with more than one painting which has the same name.) 3) Create Resources based on the previous table 4) Create an identity attribute Experience (Integer) 5) Create the following authorization policies a) Allow employees with over 20 years experience to access all paintings b) Allow employees with 10 – 20 years of experience to access painting which cost less than 50 c) Deny access to all paintings for employees with less than 10 year of experience OES Authorization Configuration We do need to create 2 authorization policies with specific conditions a) Allow employees with over 20 years experience to access all paintings b) Allow employees with 10 – 20 years of experience to access painting which cost less than 50 c) Deny access to all paintings for employees with less than 10 year of experience We don’t need an explicit policy for Deny access to all paintings for employees with less than 10 year of experience, because Oracle Entitlements Server will automatically deny if there is no matching policy. OEG Policy The OEG policy looks like the following The 11g Authorization filter configuration is similar to :  The ${PAINTING_NAME} and ${USER_EXPERIENCE} variables are initialized by the "Retrieve from the HTTP header" filters for testing purpose. That's to say, under Service Explorer, we need to provide 2 attributes "Experience" & "Painting" following the OES 11g Authorization filter described above.

    Read the article

  • Math with Timestamp

    - by Knut Vatsendvik
    table.sql { border-width: 1px; border-spacing: 2px; border-style: dashed; border-color: #0023ff; border-collapse: separate; background-color: white; } table.sql th { border-width: 1px; padding: 1px; border-style: none; border-color: gray; background-color: white; -moz-border-radius: 0px 0px 0px 0px; } table.sql td { border-width: 1px; padding: 3px; border-style: none; border-color: gray; background-color: white; -moz-border-radius: 0px 0px 0px 0px; } .sql-keyword { color: #0000cd; background-color: inherit; } .sql-result { color: #458b74; background-color: inherit; } Got this little SQL quiz from a colleague.  How to add or subtract exactly 1 second from a Timestamp?  Sounded simple enough at first blink, but was a bit trickier than expected. If the data type had been a Date, we knew that we could add or subtract days, minutes or seconds using + or – sysdate + 1 to add one day sysdate - (1 / 24) to subtract one hour sysdate + (1 / 86400) to add one second Would the same arithmetic work with Timestamp as with Date? Let’s test it out with the following query SELECT   systimestamp , systimestamp + (1 / 86400) FROM dual; ---------- 03.05.2010 22.11.50,240887 +02:00 03.05.2010 The first result line shows us the system time down to fractions of seconds. The second result line shows the result as Date (as used for date calculation) meaning now that the granularity is reduced down to a second.   By using the PL/SQL dump() function, we can confirm this with the following query SELECT   dump(systimestamp) , dump(systimestamp + (1 / 86400)) FROM dual; ---------- Typ=188 Len=20: 218,7,5,4,8,53,9,0,200,46,89,20,2,0,5,0,0,0,0,0 Typ=13 Len=8: 218,7,5,4,10,53,10,0 Where typ=13 is a runtime representation for Date. So how can we increase the precision to include fractions of second? After investigating it a bit, we found out that the interval data type INTERVAL DAY TO SECOND could be used with the result of addition or subtraction being a Timestamp. Let’s try again our first query again, now using the interval data type. SELECT systimestamp,    systimestamp + INTERVAL '0 00:00:01.0' DAY TO SECOND(1) FROM dual; ---------- 03.05.2010 22.58.32,723659000 +02:00 03.05.2010 22.58.33,723659000 +02:00 Yes, it worked! To finish the story, here is one example showing how to specify an interval of 2 days, 6 hours, 30 minutes, 4 seconds and 111 thousands of a second. INTERVAL ‘2 6:30:4.111’ DAY TO SECOND(3)

    Read the article

  • HOWTO Turn off SPARC T4 or Intel AES-NI crypto acceleration.

    - by darrenm
    Since we released hardware crypto acceleration for SPARC T4 and Intel AES-NI support we have had a common question come up: 'How do I test without the hardware crypto acceleration?'. Initially this came up just for development use so developers can do unit testing on a machine that has hardware offload but still cover the code paths for a machine that doesn't (our integration and release testing would run on all supported types of hardware anyway).  I've also seen it asked in a customer context too so that we can show that there is a performance gain from the hardware crypto acceleration, (not just the fact that SPARC T4 much faster performing processor than T3) and measure what it is for their application. With SPARC T2/T3 we could easily disable the hardware crypto offload by running 'cryptoadm disable provider=n2cp/0'.  We can't do that with SPARC T4 or with Intel AES-NI because in both of those classes of processor the encryption doesn't require a device driver instead it is unprivileged user land callable instructions. Turns out there is away to do this by using features of the Solaris runtime loader (ld.so.1). First I need to expose a little bit of implementation detail about how the Solaris Cryptographic Framework is implemented in Solaris 11.  One of the new Solaris 11 features of the linker/loader is the ability to have a single ELF object that has multiple different implementations of the same functions that are selected at runtime based on the capabilities of the machine.  The alternate to this is having the application coded to call getisax() and make the choice itself.  We use this functionality of the linker/loader when we build the userland libraries for the Solaris Cryptographic Framework (specifically libmd.so, and the unfortunately misnamed due to historical reasons libsoftcrypto.so) The Solaris linker/loader allows control of a lot of its functionality via environment variables, we can use that to control the version of the cryptographic functions we run.  To do this we simply export the LD_HWCAP environment variable with values that tell ld.so.1 to not select the HWCAP section matching certain features even if isainfo says they are present.  For SPARC T4 that would be: export LD_HWCAP="-aes -des -md5 -sha256 -sha512 -mont -mpul" and for Intel systems with AES-NI support: export LD_HWCAP="-aes" This will work for consumers of the Solaris Cryptographic Framework that use the Solaris PKCS#11 libraries or use libmd.so interfaces directly.  It also works for the Oracle DB and Java JCE.  However does not work for the default enabled OpenSSL "t4" or "aes-ni" engines (unfortunately) because they do explicit calls to getisax() themselves rather than using multiple ELF cap sections. However we can still use OpenSSL to demonstrate this by explicitly selecting "pkcs11" engine  using only a single process and thread.  $ openssl speed -engine pkcs11 -evp aes-128-cbc ... type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 54170.81k 187416.00k 489725.70k 805445.63k 1018880.00k $ LD_HWCAP="-aes" openssl speed -engine pkcs11 -evp aes-128-cbc ... type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 29376.37k 58328.13k 79031.55k 86738.26k 89191.77k We can clearly see the difference this makes in the case where AES offload to the SPARC T4 was disabled. The "t4" engine is faster than the pkcs11 one because there is less overhead (again on a SPARC T4-1 using only a single process/thread - using -multi you will get even bigger numbers). $ openssl speed -evp aes-128-cbc ... type 16 bytes 64 bytes 256 bytes 1024 bytes 8192 bytes aes-128-cbc 85526.61k 89298.84k 91970.30k 92662.78k 92842.67k Yet another cool feature of the Solaris linker/loader, thanks Rod and Ali. Note these above openssl speed output is not intended to show the actual performance of any particular benchmark just that there is a significant improvement from using hardware acceleration on SPARC T4. For cryptographic performance benchmarks see the http://blogs.oracle.com/BestPerf/ postings.

    Read the article

  • Overriding the Pager rendering in Orchard

    - by Bertrand Le Roy
    The Pager shape that is used in Orchard to render pagination is one of those shapes that are built in code rather than in a Razor template. This can make it a little more confusing to override, but nothing is impossible. If we look at the Pager method in CoreShapes, here is what we see: [Shape] public IHtmlString Pager(dynamic Shape, dynamic Display) { Shape.Metadata.Alternates.Clear(); Shape.Metadata.Type = "Pager_Links"; return Display(Shape); } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The Shape attribute signals a shape method. All it does is remove all alternates that may exist and replace the type of the shape with “Pager_Links”. In turn, this shape method is rather large and complicated, but it renders as a set of smaller shapes: a List with a “pager” class, and under that Pager_First, Pager_Previous, Pager_Gap, for each page a Pager_Link or a Pager_Current, then Pager_Gap, Pager_Next and Pager_Last. Each of these shapes can be displayed or not depending on the properties of the pager. Each can also be overridden with a Razor template. This can be done by dropping a file into the Views folder of your theme. For example, if you want the current page to appear between square braces, you could drop this Pager-CurrentPage.cshtml into your views folder: <span>[@Model.Value]</span> This overrides the original shape method, which was this: [Shape] public IHtmlString Pager_CurrentPage(HtmlHelper Html, dynamic Display, object Value) { var tagBuilder = new TagBuilder("span"); tagBuilder.InnerHtml = Html.Encode(Value is string ? (string)Value : Display(Value)); return MvcHtmlString.Create(tagBuilder.ToString()); } And here is what it would look like: Now what if we want to completely hide the pager if there is only one page? Well, the easiest way to do that is to override the Pager shape by dropping the following into the Views folder of your theme: @{ if (Model.TotalItemCount > Model.PageSize) { Model.Metadata.Alternates.Clear(); Model.Metadata.Type = "Pager_Links"; @Display(Model) } } And that’s it. The code in this template just adds a check for the number of items to display (in a template, Model is the shape) and only displays the Pager_Links shape if it knows that there’s going to be more than one page.

    Read the article

  • Green (Screen) Computing

    - by onefloridacoder
    I recently was given an assignment to create a UX where a user could use the up and down arrow keys, as well as the tab and enter keys to move through a Silverlight datagrid that is going be used as part of a high throughput data entry UI. And to be honest, I’ve not trapped key codes since I coded JavaScript a few years ago.  Although the frameworks I’m using made it easy, it wasn’t without some trial and error.    The other thing that bothered me was that the customer tossed this into the use case as they were delivering the use case.  Fine.  I’ll take a whack at anything and beat up myself and beg (I’m not beyond begging for help) the community for help to get something done if I have to. It wasn’t as bad as I thought and I thought I would hopefully save someone a few keystrokes if you wanted to build a green screen for your customer.   Here’s the ValueConverter to handle changing the strings to decimals and then back again.  The value is a nullable valuetype so there are few extra steps to take.  Usually the “ConvertBack()” method doesn’t get addressed but in this case we have two-way binding and the converter needs to ensure that if the user doesn’t enter a value it will remain null when the value is reapplied to the model object’s setter.  1: using System; 2: using System.Windows.Data; 3: using System.Globalization; 4:  5: public class NullableDecimalToStringConverter : IValueConverter 6: { 7: public object Convert(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture) 8: { 9: if (!(((decimal?)value).HasValue)) 10: { 11: return (decimal?)null; 12: } 13: if (!(value is decimal)) 14: { 15: throw new ArgumentException("The value must be of type decimal"); 16: } 17:  18: NumberFormatInfo nfi = culture.NumberFormat; 19: nfi.NumberDecimalDigits = 4; 20:  21: return ((decimal)value).ToString("N", nfi); 22: } 23:  24: public object ConvertBack(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture) 25: { 26: decimal nullableDecimal; 27: decimal.TryParse(value.ToString(), out nullableDecimal); 28:  29: return nullableDecimal == 0 ? null : nullableDecimal.ToString(); 30: } 31: }            The ConvertBack() method uses TryParse to create a value from the incoming string so if the parse fails, we get a null value back, which is what we would expect.  But while I was testing I realized that if the user types something like “2..4” instead of “2.4”, TryParse will fail and still return a null.  The user is getting “puuu-lenty” of eye-candy to ensure they know how many values are affected in this particular view. Here’s the XAML code.   This is the simple part, we just have a DataGrid with one column here that’s bound to the the appropriate ViewModel property with the Converter referenced as well. 1: <data:DataGridTextColumn 2: Header="On-Hand" 3: Binding="{Binding Quantity, 4: Mode=TwoWay, 5: Converter={StaticResource DecimalToStringConverter}}" 6: IsReadOnly="False" /> Nothing too magical here.  Just some XAML to hook things up.   Here’s the code behind that’s handling the DataGridKeyup event.  These are wired to a local/private method but could be converted to something the ViewModel could use, but I just need to get this working for now. 1: // Wire up happens in the constructor 2: this.PicDataGrid.KeyUp += (s, e) => this.HandleKeyUp(e);   1: // DataGrid.BeginEdit fires when DataGrid.KeyUp fires. 2: private void HandleKeyUp(KeyEventArgs args) 3: { 4: if (args.Key == Key.Down || 5: args.Key == Key.Up || 6: args.Key == Key.Tab || 7: args.Key == Key.Enter ) 8: { 9: this.PicDataGrid.BeginEdit(); 10: } 11: }   And that’s it.  The ValueConverter was the biggest problem starting out because I was using an existing converter that didn’t take nullable value types into account.   Once the converter was passing back the appropriate value (null, “#.####”) the grid cell(s) and the model objects started working as I needed them to. HTH.

    Read the article

  • Controlling server configurations with IPS

    - by barts
    I recently received a customer question regarding how they best could control which packages and which versions were used on their production Solaris 11 servers.  They had considered pointing each server at its own software repository - a common initial approach.  A simpler method leverages one of dependency mechanisms we introduced with Solaris 11, but is not immediately obvious to most people. Typically, most internal IT departments qualify particular versions for production use.  What this customer wanted to do was insure that their operations staff only installed internally qualified versions of Solaris on their servers.  The easiest way of doing this is to leverage the 'incorporate' type of dependency in a small package defined for each server type.  From the reference " Packaging and Delivering Software With the Image Packaging System in Oracle® Solaris 11.1":  The incorporate dependency specifies that if the given package is installed, it must be at the given version, to the given version accuracy. For example, if the dependent FMRI has a version of 1.4.3, then no version less than 1.4.3 or greater than or equal to 1.4.4 satisfies the dependency. Version 1.4.3.7 does satisfy this example dependency. The common way to use incorporate dependencies is to put many of them in the same package to define a surface in the package version space that is compatible. Packages that contain such sets of incorporate dependencies are often called incorporations. Incorporations are typically used to define sets of software packages that are built together and are not separately versioned. The incorporate dependency is heavily used in Oracle Solaris to ensurethat compatible versions of software are installed together. An example incorporate dependency is: depend type=incorporate fmri=pkg:/driver/network/ethernet/[email protected],5.11-0.175.0.0.0.2.1 So, to make sure only qualified versions are installed on a server, create a package that will be installed on the machines to be controlled.  This package will contain an incorporate dependency on the "entire" package, which controls the various components used to be build Solaris.  Every time a new version of Solaris has been qualified for production use, create a new version of this package specifying the new version of "entire" that was qualified.  Once this new control package is available in the repositories configured on the production server, the pkg update command will update that system to the specified version.  Unless a new version of the control package is made available, pkg update will report that no updates are available since no version of the control package can be installed that satisfies the incorporate constraint. Note that if desired, the same package can be used to specify which packages must be present on the system by adding either "require" or "group" dependencies; the latter permits removal of some of the packages, the former does not.  More details on this can be found in either the section 5 pkg man page or the previously mentioned reference document. This technique of using package dependencies to constrain system configuration leverages the SAT solver which is at the heart of IPS, and is basic to how we package Solaris itself.  

    Read the article

  • Facebook: Sending private messages to FB profile from a static website [migrated]

    - by Frondor
    I need to setup a static website for people to: Complete a form. And using anything from Facebook API, GET the form output via message to a Facebook Profile. I've been punching my head against "facebook developers" page all night long and can't find out how to do it. Seems quite easy, but the problem is that I don't know if you'll get my point :) Like the Send Dialog feature, you can set a certain user as recipient which will be displayed on the "To:" field once the dialog appears. FB.ui({ method: 'send', to: 'UserID', link: 'http://www.nytimes.com/2011/06/15/arts/people-argue-just-to-win-scholars-assert.html', }); Ok, All I need is to be able to use the same behavior but instead of setting a "to:" parameter, I'd like to set a "message:" parameter. I don't know how I can solve this becuase there's no parameter like this on the API actually. This is what I need to build (It's a prototype, this code won't work) <form action="mysite.com" id="order"> <input type="radio" name="chocolate" value="white">White <br/> <input type="radio" name="chocolate" value="black">Black <br/> <input type="submit" value="Order" /> </form> jQuery gets the values $(document).ready(function() { $("#order").on("submit", function(e) { e.preventDefault(); var formOutput = $(this).serialize(); var order = "I'd like to eat" + formOutput + "chocolate"; }); }); Facebook sdk sends this output ('order' string) FB.ui({ method: 'send', //or whatever to: 'UserID', message: order, //Its just an example, note the variable coming from the form link: 'http://www.nytimes.com/2011/06/15/arts/people-argue-just-to-win-scholars-assert.html', }); As we all know, what I wrote isn't possible, so I'm asking for any alternative solution if somebody can give me, I'm not very friendly with facebook APIs :) I though in another solution which consist in using the form output directly on the 'link:' parameter of FB.ui and then reading it with jQuery on some landing page. For example, on the message sent, the linked content redirects to this URL: http://mysite.com/dashboard.html?chocolate=white and the dashboard page source code: <script> var choco = getUrlParameter('chocolate'); $("#dashboard").text("This person wants" + choco + "chocolate") </script> <div id="dashboard"></div> And this way, I will be able to see which kind of chocolate the person selected by parsing some parameters on the URL when clicking on the link section of the message: using a code like this: FB.ui({ method: 'send', //or whatever to: 'MyUserID', link: 'http://mysite.com/dashboard.html?chocolate=white', }); But no this try, my biggest problem is that I don't know how to dynamically "customize" that "link:" paramenter with jQuery. I think the best solution is to use a code like this along with the dashboard page in order to "translate" the shared URLs and see what kind of chocolate people are demanding xD FB.ui({ //declaring a variable (example) var string = getFormData().serialize; var orderString = "mysite.com/dashboard.html?" + string; // end the variables // start facebook API code method: 'send', //or whatever to: 'MyUserID', link: orderString, }); I was working here until I gave up and started to post this http://jsfiddle.net/Frondor/sctepn06/2/ Thanks in advance, I'll love you for ever if you help me solving this :D

    Read the article

  • VirtualBox Clone Root HD / Ubuntu / Network issue

    - by john.graves(at)oracle.com
    When you clone a root Ubuntu disk in VirtualBox, one thing that gets messed up is the network card definition.  This is because Ubuntu (as it should) uses UDEV IDs for the network device.  When you boot your new disk, the network device ID has changed, so it creates a new eth1 device.  Unfortunately, this conflicts with the VirtualBox network setup.  What to do? Boot the box (no network) Edit the /etc/udev/rules.d/70-persistent-net.rules Delete the eth0 line and modify the eth1 line to be eth0 --------- Example OLD ----------- # This file was automatically generated by the /lib/udev/write_net_rules # program, run by the persistent-net-generator.rules rules file. # # You can modify it, as long as you keep each rule on a single # line, and change only the value of the NAME= key. # PCI device 0x8086:0x100e (e1000) <-------------------- Delete these two lines SUBSYSTEM=="net", ACTION=="add", DRIVERS=="?*", ATTR{address}=="08:00:27:d8:8d:15", ATTR{dev_id}=="0x0", ATTR{type}=="1", KERNEL=="eth*", NAME="eth0" # PCI device 0x8086:0x100e (e1000) ---Modify the next line and change eth1 to be eth0 SUBSYSTEM=="net", ACTION=="add", DRIVERS=="?*", ATTR{address}=="08:00:27:89:84:98", ATTR{dev_id}=="0x0", ATTR{type}=="1", KERNEL=="eth*", NAME="eth1" .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } ---------------------------------------- --------- Example NEW ----------- # This file was automatically generated by the /lib/udev/write_net_rules # program, run by the persistent-net-generator.rules rules file. # # You can modify it, as long as you keep each rule on a single # line, and change only the value of the NAME= key. # PCI device 0x8086:0x100e (e1000) SUBSYSTEM=="net", ACTION=="add", DRIVERS=="?*", ATTR{address}=="08:00:27:89:84:98", ATTR{dev_id}=="0x0", ATTR{type}=="1", KERNEL=="eth*", NAME="eth0" .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } ----------------------------------------

    Read the article

  • RewriteRule not working for certain URLs

    - by keiki
    There are a few domains pointing towards the same server, and of course I need them all redirect to only one of them. Redirects work, but only for certain URLs. What works: http://www.domain.com, http://domain.com, domain.com/index.html, domain.com/index.php, , domain.com/nonExistentDirectory, and if I click in the menu the following URLs are also redirected correctly: domain.com/foo/bar, domain.com/foo/bar.html or .php or other extension. What doesn't work: domain.com/existentDirectory, domain.com/foo/bar (if I type the URL in the address bar). If anyone will have the time and skill and will to tell me where's the mistake, I'll be deeply grateful. Here's my .htaccess file: AddHandler x-httpd-php .html .htm <ifModule mod_gzip.c> mod_gzip_on Yes mod_gzip_dechunk Yes mod_gzip_item_include file \.(html?|txt|css|js|php|pl)$ mod_gzip_item_include handler ^cgi-script$ mod_gzip_item_include mime ^text/.* mod_gzip_item_include mime ^application/x-javascript.* mod_gzip_item_exclude mime ^image/.* mod_gzip_item_exclude rspheader ^Content-Encoding:.*gzip.* </ifModule> <ifModule mod_expires.c> ExpiresActive On ExpiresDefault "access plus 1 seconds" ExpiresByType text/html "access plus 1 seconds" ExpiresByType image/gif "access plus 2592000 seconds" ExpiresByType image/jpeg "access plus 2592000 seconds" ExpiresByType image/png "access plus 2592000 seconds" ExpiresByType text/css "access plus 2592000 seconds" ExpiresByType text/javascript "access plus 216000 seconds" ExpiresByType application/x-javascript "access plus 216000 seconds" </ifModule> <ifModule mod_headers.c> <filesMatch "\\.(ico|pdf|flv|jpg|jpeg|png|gif|swf)$"> Header set Cache-Control "max-age=2592000, public" </filesMatch> <filesMatch "\\.(css)$"> Header set Cache-Control "max-age=2592000, public" </filesMatch> <filesMatch "\\.(js)$"> Header set Cache-Control "max-age=216000, private" </filesMatch> <filesMatch "\\.(xml|txt)$"> Header set Cache-Control "max-age=216000, public, must-revalidate" </filesMatch> <filesMatch "\\.(html|htm|php)$"> Header set Cache-Control "max-age=1, private, must-revalidate" </filesMatch> </ifModule> <ifModule mod_headers.c> Header unset ETag </ifModule> FileETag None <ifModule mod_headers.c> Header unset Last-Modified </ifModule> # BEGIN WordPress <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /index.php [L] </IfModule> # END WordPress RewriteCond %{HTTP_HOST} ^foo\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo1\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo1\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo2\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo2\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo3\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo3\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo8\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo8\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] Thinking that the above version was overkill, I've also tried to redirect all the requests for domains different than the main on to be redirected to it like this: RewriteCond %{HTTP_HOST} !^domain\.com$ [NC] RewriteRule ^(.*)$ http://domain.com [L,R=301] Is it also wrong? Because it doesn't work either! P.S. @Sodved I've tried that and it doesn't help (I comment here because I can't seem to be able to comment your answer.) Removing the following piece of code didn't solve the issue either, so the problem must be somewhere else: # BEGIN WordPress <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /index.php [L] </IfModule> # END WordPress New details: using this tool for checking the redirects I got the following results for the URLs that are not redirected: Checked link: http://domain.com/aDirectory/ Type of link: direct link (note the trailing slash above) and: Checked link: http://domain.com/aDirectory Type of redirect: 301 Moved Permanently Redirected to: http://domain.com/aDirectory/ (no trailing slash here) I hope/suspect I'm getting closer to the cause of this behavior.

    Read the article

  • Google Fetch issue

    - by Karen
    When I do a Google fetch on any of my webpages the results are all the same (below). I'm not a programmer but I'm pretty sure this is not correct. Out of all the fetches I have done only one was different and the content length was 6x below and showed meta tags etc. Maybe this explains other issues I've been having with the site: a drop in indexed pages. Meta tag analyzer says I have no title tag, meta tags or description even though I do it on all pages. I had an SEO team working on the site and they were stumped by why pages were not getting indexed. So they figure it was some type of code error. Are they right? HTTP/1.1 200 OK Cache-Control: private Content-Type: text/html; charset=utf-8 Content-Encoding: gzip Vary: Accept-Encoding Server: Microsoft-IIS/7.5 X-AspNet-Version: 4.0.30319 X-Powered-By: ASP.NET Date: Thu, 11 Oct 2012 11:45:41 GMT Content-Length: 1054 <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head> <title></title> <script type="text/javascript"> function getCookie(cookieName) { if (document.cookie.length > 0) { cookieStart = document.cookie.indexOf(cookieName + "="); if (cookieStart != -1) { cookieStart = cookieStart + cookieName.length + 1; cookieEnd = document.cookie.indexOf(";", cookieStart); if (cookieEnd == -1) cookieEnd = document.cookie.length; return unescape(document.cookie.substring(cookieStart, cookieEnd)); } } return ""; } function setTimezone() { var rightNow = new Date(); var jan1 = new Date(rightNow.getFullYear(), 0, 1, 0, 0, 0, 0); // jan 1st var june1 = new Date(rightNow.getFullYear(), 6, 1, 0, 0, 0, 0); // june 1st var temp = jan1.toGMTString(); var jan2 = new Date(temp.substring(0, temp.lastIndexOf(" ") - 1)); temp = june1.toGMTString(); var june2 = new Date(temp.substring(0, temp.lastIndexOf(" ") - 1)); var std_time_offset = (jan1 - jan2) / (1000 * 60 * 60); var daylight_time_offset = (june1 - june2) / (1000 * 60 * 60); var dst; if (std_time_offset == daylight_time_offset) { dst = "0"; // daylight savings time is NOT observed } else { // positive is southern, negative is northern hemisphere var hemisphere = std_time_offset - daylight_time_offset; if (hemisphere >= 0) std_time_offset = daylight_time_offset; dst = "1"; // daylight savings time is observed } var exdate = new Date(); var expiredays = 1; exdate.setDate(exdate.getDate() + expiredays); document.cookie = "TimeZoneOffset=" + std_time_offset + ";"; document.cookie = "Dst=" + dst + ";expires=" + exdate.toUTCString(); } function checkCookie() { var timeOffset = getCookie("TimeZoneOffset"); var dst = getCookie("Dst"); if (!timeOffset || !dst) { setTimezone(); window.location.reload(); } } </script> </head> <body onload="checkCookie()"> </body> </html>

    Read the article

  • How do I make a firefox extension execute script on page open/load? [migrated]

    - by Will Mc
    Thanks in advance! I am creating my first extension (A firefox extension). See below for full description of final product. I need help starting off. I have looked and studied the HelloWorld.xpi example found on Mozilla's site so I am happy to edit that to learn. In the example, when you click a menu item it runs script to display an alert message. My question is, how would I edit this extension to run the script on page load? I am guessing I need to insert some code in the browserOverlay as it loads on page load so, here is the browserOverlay.xpi from the example I am editing to learn: <?xml version="1.0"?> <?xml-stylesheet type="text/css" href="chrome://global/skin/" ?> <?xml-stylesheet type="text/css" href="chrome://xulschoolhello/skin/browserOverlay.css" ?> <!DOCTYPE overlay SYSTEM "chrome://xulschoolhello/locale/browserOverlay.dtd"> <overlay id="xulschoolhello-browser-overlay" xmlns="http://www.mozilla.org/keymaster/gatekeeper/there.is.only.xul"> <script type="application/x-javascript" src="chrome://xulschoolhello/content/browserOverlay.js" /> <stringbundleset id="stringbundleset"> <stringbundle id="xulschoolhello-string-bundle" src="chrome://xulschoolhello/locale/browserOverlay.properties" /> </stringbundleset> <menubar id="main-menubar"> <menu id="xulschoolhello-hello-menu" label="&xulschoolhello.hello.label;" accesskey="&xulschoolhello.helloMenu.accesskey;" insertafter="helpMenu"> <menupopup> <menuitem id="xulschoolhello-hello-menu-item" label="&xulschoolhello.hello.label;" accesskey="&xulschoolhello.helloItem.accesskey;" oncommand="XULSchoolChrome.BrowserOverlay.sayHello(event);" /> </menupopup> </menu> </menubar> <vbox id="appmenuSecondaryPane"> <menu id="xulschoolhello-hello-menu-2" label="&xulschoolhello.hello.label;" accesskey="&xulschoolhello.helloMenu.accesskey;" insertafter="appmenu_addons"> <menupopup> <menuitem id="xulschoolhello-hello-menu-item-2" label="&xulschoolhello.hello.label;" accesskey="&xulschoolhello.helloItem.accesskey;" oncommand="XULSchoolChrome.BrowserOverlay.sayHello(event);" /> </menupopup> </menu> </vbox> </overlay> I hope you can help me. I need to know what code I should use and where I should put it... Here is the gist of my overall extension - I am creating a click to call extension. This extension will search any new page for a phone number whether just refreshed, new page, new tab etc... Each phone number when clicked will open a new tab and direct user to a custom URL. Thanks again!

    Read the article

  • Need help partitioning when reinstalling Ubuntu 14.04

    - by Chris M.
    I upgraded to 14.04 about a month ago on my HP Mini netbook (about 16 GB hard disk). A few days ago the system crashed (I don't know why but I was using internet at the time). When I restarted the computer, Ubuntu would not load. Instead, I got a message from the BIOS saying Reboot and Select proper Boot device or Insert Boot Media in selected Boot device and press a key I took this to mean that I needed to reinstall 14.04. When I try to reinstall Ubuntu from the USB stick, I choose "Erase disk and install Ubuntu" but then I get a message: Some of the partitions you created are too small. Please make the following partitions at least this large: / 3.3 GB If you do not go back to the partitioner and increase the size of these partitions, the installation may fail. At first I hit Continue to see if it would install anyway, and it gave the message: The attempt to mount a file system with type ext4 in SCSI1 (0,0,0), partition # 1 (sda) at / failed. You may resume partitioning from the partitioning menu. The second time I hit Go Back, and it took me to the following partitioning table: Device Type Mount Point Format Size Used System /dev/sda /dev/sda1 ext4 (checked) 3228 MB Unknown /dev/sda5 swap (not checked) 1063 MB Unknown + - Change New Partition Table... Revert Device for boot loader installation: /dev/sda ATA JM Loader 001 (4.3 GB) At this point I'm not sure what to do. I've never partitioned my hard drive before and I don't want to screw things up. (I'm not particularly tech savvy.) Can you instruct me what I should do. (P.S. I'm afraid the table might not appear as I typed it in.) Results from fdisk: ubuntu@ubuntu:~$ sudo fdisk -l Disk /dev/sda: 4294 MB, 4294967296 bytes 255 heads, 63 sectors/track, 522 cylinders, total 8388608 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00000000 Disk /dev/sda doesn't contain a valid partition table Disk /dev/sdb: 7860 MB, 7860125696 bytes 155 heads, 31 sectors/track, 3194 cylinders, total 15351808 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x0009a565 Device Boot Start End Blocks Id System /dev/sdb1 * 2768 15351807 7674520 b W95 FAT32 ubuntu@ubuntu:~$ Here is what it displays when I open the Disks utility (I tried the screenshot terminal command you suggested but it didn't seem to do anything): 4.3 GB Hard Disk /dev/sda Model: JM Loader 001 (01000001) Size: 4.3 GB (4,294,967,296 bytes) Serial Number: 01234123412341234 Assessment: SMART is not supported Volumes Size: 4.3 GB (4,294,967,296 bytes) Device: /dev/sda Contents: Unknown (There is a button in the utility that when you click it gives the following options: Format... Create Disk Image... Restore Disk Image... Benchmark but SMART Data & Self-Tests... is dimmed out) When I hit F9 Change Boot Device Order, it shows the hard drive as: SATA:PM-JM Loader 001 When I hit F10 to get me into the BIOS Setup Utility, under Diagnostic it shows: Primary Hard Disk Self Test Not Support NetworkManager Tool State: disconnected Device: eth0 Type: Wired Driver: atl1c State: unavailable Default: no HW Address: 00:26:55:B0:7F:0C Capabilities: Carrier Detect: yes Wired Properties Carrier: off When I run command lshw -C network, I get: WARNING: you should run this program as super-user. *-network description: Network controller product: BCM4312 802.11b/g LP-PHY vendor: Broadcom Corporation physical id: 0 bus info: pci@0000:01:00.0 version: 01 width: 64 bits clock: 33MHz capabilities: bus_master cap_list configuration: driver=b43-pci-bridge latency=0 resources: irq:16 memory:feafc000-feafffff *-network description: Ethernet interface product: AR8132 Fast Ethernet vendor: Qualcomm Atheros physical id: 0 bus info: pci@0000:02:00.0 logical name: eth0 version: c0 serial: 00:26:55:b0:7f:0c capacity: 100Mbit/s width: 64 bits clock: 33MHz capabilities: bus_master cap_list ethernet physical tp 10bt 10bt-fd 100bt 100bt-fd autonegotiation configuration: autonegotiation=on broadcast=yes driver=atl1c driverversion=1.0.1.1-NAPI latency=0 link=no multicast=yes port=twisted pair resources: irq:43 memory:febc0000-febfffff ioport:ec80(size=128) WARNING: output may be incomplete or inaccurate, you should run this program as super-user.

    Read the article

  • Strings in .NET are Enumerable

    - by Scott Dorman
    It seems like there is always some confusion concerning strings in .NET. This is both from developers who are new to the Framework and those that have been working with it for quite some time. Strings in the .NET Framework are represented by the System.String class, which encapsulates the data manipulation, sorting, and searching methods you most commonly perform on string data. In the .NET Framework, you can use System.String (which is the actual type name or the language alias (for C#, string). They are equivalent so use whichever naming convention you prefer but be consistent. Common usage (and my preference) is to use the language alias (string) when referring to the data type and String (the actual type name) when accessing the static members of the class. Many mainstream programming languages (like C and C++) treat strings as a null terminated array of characters. The .NET Framework, however, treats strings as an immutable sequence of Unicode characters which cannot be modified after it has been created. Because strings are immutable, all operations which modify the string contents are actually creating new string instances and returning those. They never modify the original string data. There is one important word in the preceding paragraph which many people tend to miss: sequence. In .NET, strings are treated as a sequence…in fact, they are treated as an enumerable sequence. This can be verified if you look at the class declaration for System.String, as seen below: // Summary:// Represents text as a series of Unicode characters.public sealed class String : IEnumerable, IComparable, IComparable<string>, IEquatable<string> The first interface that String implements is IEnumerable, which has the following definition: // Summary:// Exposes the enumerator, which supports a simple iteration over a non-generic// collection.public interface IEnumerable{ // Summary: // Returns an enumerator that iterates through a collection. // // Returns: // An System.Collections.IEnumerator object that can be used to iterate through // the collection. IEnumerator GetEnumerator();} As a side note, System.Array also implements IEnumerable. Why is that important to know? Simply put, it means that any operation you can perform on an array can also be performed on a string. This allows you to write code such as the following: string s = "The quick brown fox";foreach (var c in s){ System.Diagnostics.Debug.WriteLine(c);}for (int i = 0; i < s.Length; i++){ System.Diagnostics.Debug.WriteLine(s[i]);} If you executed those lines of code in a running application, you would see the following output in the Visual Studio Output window: In the case of a string, these enumerable or array operations return a char (System.Char) rather than a string. That might lead you to believe that you can get around the string immutability restriction by simply treating strings as an array and assigning a new character to a specific index location inside the string, like this: string s = "The quick brown fox";s[2] = 'a';   However, if you were to write such code, the compiler will promptly tell you that you can’t do it: This preserves the notion that strings are immutable and cannot be changed once they are created. (Incidentally, there is no built in way to replace a single character like this. It can be done but it would require converting the string to a character array, changing the appropriate indexed location, and then creating a new string.)

    Read the article

  • WF4 &ndash; Guess the number game!

    - by MarkPearl
    I posted yesterday how really good WF4 was looking. Today I thought I would show some real basics that I was able to figure out. This will be a simple example, I am going to make a flowchart workflow – which will prompt the user to guess the number until they guess the right number. Lets begin… Make a new project and make it a Workflow console Application. Then select the Workflow file and drag a FlowChart (2) to point 3. This will now show a green start circle in the designer form. We are going to work with primitives to start with. We are now going to drag a few objects onto the Workflow, We drag the WriteLine, Assign & Decision items onto the designer. Once they are dragged onto the designer we will want to link them up. The order that they are linked is critical since this will determine the order of the solution. In this case, we want the system to first ask “Guess a number”, then to wait for the user to input some code, and then to display “You got it” if they got it right, and “Try again” if they got it wrong. So we now link the arrows to the objects. This is done by moving the mouse pointer over the start objects and clicking on one of the toggles and then dragging it to the next object and releasing the button over one of the toggles. This will place an arrow from the source object to the target object. Okay… pretty simple stuff – now we just need these primitive objects to do stuff. Lets start with the WriteLine primitive. We place the text in inverted commas in the Text field. Because this field accepts any valid VB expression we could have put variables etc. in there if we wanted to. The next thing we want to do is allow the user to input a number. This brings up an interesting problem, if a user were to type in a number, there would need to be someway to declare a variable to hold that value for the life of the workflow. We can achieve this by declaring a variable. To declare a variable, move your cursor over the variables tab at the bottom of the workflow, and then type the name of the new variable in the “Create Variable” field and set it as shown in the image above. Now that we have a variable, we want to call the Console.Readline method and assign the inputted value from the Console to that variable. The code that cannot be seen is actually this – Convert.ToInt32(Console.ReadLine()) We now have a workflow that first prompts the user for a number, then allows the user to type in a number. We are almost done, we just need to make the system react to the value inputted. There are a few ways we could do this, I am going to use the Decision item. So select the Decision object on the designer and then view its properties (F4 for me), and in the condition field place a condition. For simplicity sake I have decided that if the user guesses 10, they will have guessed the number. This is now the completed workflow. Its really easy to understand and shows some really powerful principles for Business applications. You can run the application and see what it does. Imagine writing business solutions that do not worry about the exact flow of objects, but simply allows a business analyst or someone to configure the solution to work exactly as the business rules would dictate. And if the rules changed six months later all they would need to do is re-drag some of the flows. Now I do not know if WF4 will allow for this, but it feels like it is a step in the right direct.

    Read the article

< Previous Page | 379 380 381 382 383 384 385 386 387 388 389 390  | Next Page >