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• Ubuntu hangs on booting up after a update

  - by alFReD NSH

  I've made a clean install yesterday, for the first time restarted, everything went good and then after I updated packages and copied my old home directory to replace the new one, when I restarted it hung when it was booting. I tried reinstalling again and doing the same thing, but again same thing happened. Here's what I see, before when the Ubuntu logo with the five dots is shown: Then after that, 3 or 4 of the dots will load and hangs there. If I press arrow up before that, this will be shown I started my laptop again today(the pictures are for the day before) and after that, boot up with live CD and got the logs. dmesg: http://pastebin.com/aVxV7BQF syslog: http://pastebin.com/4E2BrRUK And some info: alfred@alFitop:~$ uname -a Linux alFitop 3.2.0-24-generic #39-Ubuntu SMP Mon May 21 16:52:17 UTC 2012 x86_64 x86_64 x86_64 GNU/Linux lshw: http://pastebin.com/AZbKJmsT sources.list : http://pastebin.com/2HazmuyV My problem is a bit similar to here: http://ubuntuforums.org/showthread.php?t=1918271 Though I didn't change my x.org config. Only changed home directory and updated packages. I've tried memtest and fschk, both passed. In the recovery mode boot option, I've also realized that same things happen in failsafe graphical mode. But when I go into the network mode, I can boot up my system, but of course same the graphics are just basic. Adding blacklist intel_ips to /etc/modprobe.d/blacklist.conf solves the first message, but still I get the broken pipe and CPU stack traces. The current kernel version is 3.2.0-25, I've tried booting up in the 3.2.0-23(the one the installer came with, but same results.) Also uninstalled apparmor, didn't help. I've installed Ubuntu again, this time without copying the home directory, also same result. --- UPDATE --- This problem was solved before with removing backports, but its back again! I've updated my laptop last night and the problem came back. It's definitely one of these packages. My /var/log/apt/term.log and /var/log/apt/history.log. I'm almost having the same situation. --- UPDATE --- I realized this also have happened on times that I have updated(haven't restarted after it) and my computer power has been cut off and its shutdown due to lack of power. And I realized if I just do as I answered but not in somewhere without GUI(networking mode has the GUI) it wouldn't work.

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• What location to put bootloader, when running multiple drives and partition

  - by Matt G

  I have Win8 on my desktop, where a 120G SSD is used to run windows and some select applications, while I have a 2TB HDD to provide basic file storage and where possible, install applications instead of on the SSD. I want to install Ubuntu on a new partition of the HDD (I allocated 300GB, with 5GB swap file). I've used a USB to install the OS, which seemed to have done the job. However, after prompting for a restart, I can no longer boot to ubuntu. During instillation I was confused about where to install the "boot loader instillation". I ended up selecting "/dev/stb" because I figured I would be able to boot with BIOS by selecting the HDD drive as a priority over the SSD. The bootloader is a large part of where I think I went wrong. My partition system looked something like this: /dev/sta ... //SSD ~120 GB /dev/sta1 NTFS (350 MB) //Win8System /dev/sta2 NTFS (118 GB) //Win8C-Drive /dev/stb ... //HDD ~2TB /dev/stb1 NTFS (1563 GB) //FileStorage /dev/stb5 Free Space (300 GB) //Space I want to use for Linux (NOTE: Created two partitions from the 300GB, ~5GB and 295GB. stb5,stb6.) It'd be great if I could get an explanation of what drive you'd select for the boot loader and why, and what selections won't work with regards to the Boot Loader Instillation. I think I understand what Grub is, but I have no idea on how to use it, or play around with it. I seem to be able to get back into OS from my usb, however I believe it's just showing me a preview/trial of Ubuntu (ie, can't access any of the system NTFS drives). Note, if I try to install from the USB again, it will recognize that a version of Ubuntu 13.10 exists on the system. Apologies in advance, have used windows all my life, don't really know to much about Linux at all. Did have a brief skim over some similar questions, didn't find anything too useful. - Where to install bootloader when installing Ubuntu as secondary OS? - ubuntu 12.10 dual boot with windows 8 on two hdds - Dual-boot Windows 7 and Ubuntu on two SSDs with UEFI

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• Perl syntax error [closed]

  - by Linny

  I am a beginner taking a Perl programming course. We are trying to write a basic program for counting nucleotides in a DNA string. I'm getting syntax errors on the lines that have a single bracket on lines 28 & 70 and don't know why. It also reads that I have compilation errors. I have no idea where to start figuring that out. # The purpose of this program is to count the number of nucleotides in a strand. Each protein is counted separately # print "/n NOTE: Nucleotide counting /n"; # use strict; # enforce variable declarations use warnings; # enable compiler warnings # Display number of A,a,T,t,G,g,C,c, nucleotides in a word or sequence of letters. # my ($base) = ''; # an extracted letter from a string my ($nuceotide_count) = 0 ; # the current position within the word my ($position) = 0 ; # number of vowels in user-supplied word my ($word) = ''; # word to be processed my ($A_count) = 0 ; # of A nucleotides in the user-supplied sequence my ($a_count) = 0 ; # of A nucleotides in the user-supplied sequence my ($C_count) = 0 ; # of C nucleotides in the user-supplied sequence my ($c_count) = 0 ; # of C nucleotides in the user-supplied sequence my ($G_count) = 0 ; # of G nucleotides in the user-supplied sequence my ($g_count) = 0 ; # of G nucleotides in the user-supplied sequence my ($T_count) = 0 ; # of T nucleotides in the user-supplied sequence my ($t_count) = 0 ; # of T nucleotides in the user-supplied sequence word = (STDIN) for ($position = 0);($position if (($base eq 'a') or ($base eq 'A')) { ++$A_count; } # end if ++$position; if (($base eq 'T') or ($base eq 't')) { ++$T_count; } end if ++$position; if (($base eq 'G') or ($base eq 'g')) { ++$G_count; } # end if ++$position; if (($base eq 'C') or ($base eq 'c')) { ++$C_count; } # end if ++$position; } # end for # Display final results. # print " \n The number of A or a neucleotides is: $A_count"; print " \n The number of T or t neucleotides is: $T_count"; print " \n The number of G or g neucleotides is: $G_count"; print " \n The number of C or c neucleotides is: $C_count"; print " \n\n Program completed successfully. \n" ; exit ;

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• Is There a Real Advantage to Generic Repository?

  - by Sam

  Was reading through some articles on the advantages of creating Generic Repositories for a new app (example). The idea seems nice because it lets me use the same repository to do several things for several different entity types at once: IRepository repo = new EfRepository(); // Would normally pass through IOC into constructor var c1 = new Country() { Name = "United States", CountryCode = "US" }; var c2 = new Country() { Name = "Canada", CountryCode = "CA" }; var c3 = new Country() { Name = "Mexico", CountryCode = "MX" }; var p1 = new Province() { Country = c1, Name = "Alabama", Abbreviation = "AL" }; var p2 = new Province() { Country = c1, Name = "Alaska", Abbreviation = "AK" }; var p3 = new Province() { Country = c2, Name = "Alberta", Abbreviation = "AB" }; repo.Add<Country>(c1); repo.Add<Country>(c2); repo.Add<Country>(c3); repo.Add<Province>(p1); repo.Add<Province>(p2); repo.Add<Province>(p3); repo.Save(); However, the rest of the implementation of the Repository has a heavy reliance on Linq: IQueryable<T> Query(); IList<T> Find(Expression<Func<T,bool>> predicate); T Get(Expression<Func<T,bool>> predicate); T First(Expression<Func<T,bool>> predicate); //... and so on This repository pattern worked fantastic for Entity Framework, and pretty much offered a 1 to 1 mapping of the methods available on DbContext/DbSet. But given the slow uptake of Linq on other data access technologies outside of Entity Framework, what advantage does this provide over working directly with the DbContext? I attempted to write a PetaPoco version of the Repository, but PetaPoco doesn't support Linq Expressions, which makes creating a generic IRepository interface pretty much useless unless you only use it for the basic GetAll, GetById, Add, Update, Delete, and Save methods and utilize it as a base class. Then you have to create specific repositories with specialized methods to handle all the "where" clauses that I could previously pass in as a predicate. Is the Generic Repository pattern useful for anything outside of Entity Framework? If not, why would someone use it at all instead of working directly with Entity Framework? Edit: Original link doesn't reflect the pattern I was using in my sample code. Here is an (updated link).

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• How to suggest using an ORM instead of stored procedures?

  - by Wayne M

  I work at a company that only uses stored procedures for all data access, which makes it very annoying to keep our local databases in sync as every commit we have to run new procs. I have used some basic ORMs in the past and I find the experience much better and cleaner. I'd like to suggest to the development manager and rest of the team that we look into using an ORM Of some kind for future development (the rest of the team are only familiar with stored procedures and have never used anything else). The current architecture is .NET 3.5 written like .NET 1.1, with "god classes" that use a strange implementation of ActiveRecord and return untyped DataSets which are looped over in code-behind files - the classes work something like this: class Foo { public bool LoadFoo() { bool blnResult = false; if (this.FooID == 0) { throw new Exception("FooID must be set before calling this method."); } DataSet ds = // ... call to Sproc if (ds.Tables[0].Rows.Count > 0) { foo.FooName = ds.Tables[0].Rows[0]["FooName"].ToString(); // other properties set blnResult = true; } return blnResult; } } // Consumer Foo foo = new Foo(); foo.FooID = 1234; foo.LoadFoo(); // do stuff with foo... There is pretty much no application of any design patterns. There are no tests whatsoever (nobody else knows how to write unit tests, and testing is done through manually loading up the website and poking around). Looking through our database we have: 199 tables, 13 views, a whopping 926 stored procedures and 93 functions. About 30 or so tables are used for batch jobs or external things, the remainder are used in our core application. Is it even worth pursuing a different approach in this scenario? I'm talking about moving forward only since we aren't allowed to refactor the existing code since "it works" so we cannot change the existing classes to use an ORM, but I don't know how often we add brand new modules instead of adding to/fixing current modules so I'm not sure if an ORM is the right approach (too much invested in stored procedures and DataSets). If it is the right choice, how should I present the case for using one? Off the top of my head the only benefits I can think of is having cleaner code (although it might not be, since the current architecture isn't built with ORMs in mind so we would basically be jury-rigging ORMs on to future modules but the old ones would still be using the DataSets) and less hassle to have to remember what procedure scripts have been run and which need to be run, etc. but that's it, and I don't know how compelling an argument that would be. Maintainability is another concern but one that nobody except me seems to be concerned about.

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• Oracle GoldenGate: Knowledge Document Series Post #2

  - by Doug Reid

  0 false 18 pt 18 pt 0 0 false false false /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:12.0pt; font-family:"Times New Roman"; mso-ascii-font-family:Cambria; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Cambria; mso-hansi-theme-font:minor-latin;} 0 false 18 pt 18 pt 0 0 false false false /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:12.0pt; font-family:"Times New Roman"; mso-ascii-font-family:Cambria; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Cambria; mso-hansi-theme-font:minor-latin;} For our second post in this series the team would like to highlight the knowledge document “How-To: Oracle GoldenGate – Heartbeat Process to Monitor Lag and Performance”. This knowledge document outlines a procedure to reliably measure lag between source and target systems through the use of 'heartbeat' tables. The basic idea is to have a table on the source system that gets updated at a predetermined interval. In your capture processes you would capture the update from the heartbeat table. Using tokens you would add some additional information to the heartbeat record to be able to tell which extract process was capturing the update. This additional information would be used downstream to calculate the real lag time between the source and target systems for a given extract and by checking the last update time on the heartbeat at the target you could also determine if data has stopped flowing between the source and target.  Click here to view the document

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• Unable to connect to mail server via IMAP and roundcube

  - by mrhatter

  I am having trouble getting the final parts of my mail server up and working. I followed this tutorial to get everything set up on the mail server side. I have installed roundcube for webmail and configured it but it is saying "error connecting, connection refused" when attempting to connect to it using IMAP. This is thorough the "test imap" section of its installer. Also it is giving me an error message about perissions for it's log and temp folders but that's not as important as acutally getting mail to work. I have also tried connecting to the mail server using thunderbird however it cannot establish a connection either and I know my login information is correct. I know that the databases are working correctly based on the roundcube installer telling me that they have been "successfully initialized". Here are my firewall rules -A INPUT -i lo -j ACCEPT -A INPUT -m conntrack --ctstate RELATED,ESTABLISHED -j ACCEPT -A INPUT -p tcp -m tcp --dport 22 -j ACCEPT -A INPUT -p tcp -m tcp --dport 25 -j ACCEPT -A INPUT -p tcp -m tcp --dport 80 -j ACCEPT -A INPUT -p tcp -m tcp --dport 443 -j ACCEPT -A INPUT -p tcp -m tcp --dport 465 -j ACCEPT -A INPUT -p tcp -m tcp --dport 487 -j ACCEPT -A OUTPUT -p tcp -m tcp --dport 993 -j ACCEPT -A INPUT -j DROP Which I set up in iptables. I have modified them from what I used in this tutorial I'm not sure what to try next. Any help would be wonderful! I am using Ubuntu 14.04 server, apache 2.4.7, roundcube 1.0.1, and the latest versions of dovecot and postfix. The email databases are contained in mysql. I am running this on a VPS server. UPDATE: I have changed from iptables to using ufw. I have run the following commands to set up a basic firewall with ufw. ufw default deny ufw allow ssh ufw allow http ufw allow https ufw allow imap ufw allow imaps ufw allow smtp I then used telnet to check all of the mail ports. But Port 993 isnt working even though ufw says both 993 and 993/tcp are open. What am I missing?

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• Need help with xorg.conf for dual Radeon HD6450 video cards with 4 monitors

  - by Eriks Goodwin-Pfister

  I am running 64-bit Ubuntu 13.10 with Unity and have dual (2) Radeon HD6450 video cards and 4 Hanns-G HL273 monitors. Each Radeon card is driving one monitor via DVI and the other via VGA. I am running the proprietary video drivers from AMD's web site: "amd-catalyst-13.11-beta V9.4-linux-x86.x86_64.run" I tried to use "amd-catalyst-13.12-linux-x86.x86_64.run" but could not get that newer version to install. What I need help with is how to "correct" my xorg.conf file and any other needed instructions to get all four of my monitors to work as a continuous desktop that allows me to drag things from one monitor to the next, etc. When I tried to use the default open source drivers that came in Ubuntu 13.10, only three of the monitors would work. Now that I am running the proprietary ones, all four monitors come on and I can move my mouse from one end to the other--but only the right-most monitor displays my desktop and allows me to "do anything". Any time I move my mouse to any of the other three monitors (which display all-white), it turns into an "X" and does not do anything else but move. Enabling xinerama makes all four displays go all-black after login. I do have amdcccle installed, but it does not seem to have the ability to handle my particular configuration. My Current xorg.conf: Section "ServerLayout" Identifier "Basic Layout" Screen 0 "Screen1" 5760 0 Screen 1 "Screen0" 0 0 Screen 2 "Screen2" 3840 0 Screen 3 "Screen3" 1920 0 EndSection Section "Module" EndSection Section "Monitor" Identifier "0-DFP2" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" Option "PreferredMode" "1920x1080" Option "TargetRefresh" "60" Option "Position" "0 0" Option "Rotate" "normal" Option "Disable" "false" EndSection Section "Monitor" Identifier "0-CRT1" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" Option "PreferredMode" "1920x1080" Option "TargetRefresh" "60" Option "Position" "0 0" Option "Rotate" "normal" Option "Disable" "false" EndSection Section "Monitor" Identifier "1-DFP2" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" Option "PreferredMode" "1920x1080" Option "TargetRefresh" "60" Option "Position" "0 0" Option "Rotate" "normal" Option "Disable" "false" EndSection Section "Monitor" Identifier "1-CRT1" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" Option "PreferredMode" "1920x1080" Option "TargetRefresh" "60" Option "Position" "0 0" Option "Rotate" "normal" Option "Disable" "false" EndSection Section "Device" Identifier "Device0" Driver "fglrx" Option "Monitor-CRT1" "1-CRT1" BusID "PCI:1:0:0" EndSection Section "Device" Identifier "Device1" Driver "fglrx" Option "Monitor-DFP2" "0-DFP2" BusID "PCI:4:0:0" EndSection Section "Device" Identifier "Device2" Driver "fglrx" Option "Monitor-DFP2" "1-DFP2" BusID "PCI:1:0:0" Screen 1 EndSection Section "Device" Identifier "Device3" Driver "fglrx" Option "Monitor-CRT1" "0-CRT1" BusID "PCI:4:0:0" Screen 1 EndSection Section "Screen" Identifier "Screen0" Device "Device0" DefaultDepth 24 SubSection "Display" Depth 24 EndSubSection EndSection Section "Screen" Identifier "Screen1" Device "Device1" DefaultDepth 24 SubSection "Display" Depth 24 EndSubSection EndSection Section "Screen" Identifier "Screen2" Device "Device2" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection Section "Screen" Identifier "Screen3" Device "Device3" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection

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• efficient collision detection - tile based html5/javascript game

  - by Tom Burman

  Im building a basic rpg game and onto collisions/pickups etc now. Its tile based and im using html5 and javascript. i use a 2d array to create my tilemap. Im currently using a switch statement for whatever key has been pressed to move the player, inside the switch statement. I have if statements to stop the player going off the edge of the map and viewport and also if they player is about to land on a tile with tileID 3 then the player stops. Here is the statement: canvas.addEventListener('keydown', function(e) { console.log(e); var key = null; switch (e.which) { case 37: // Left if (playerX > 0) { playerX--; } if(board[playerX][playerY] == 3){ playerX++; } break; case 38: // Up if (playerY > 0) playerY--; if(board[playerX][playerY] == 3){ playerY++; } break; case 39: // Right if (playerX < worldWidth) { playerX++; } if(board[playerX][playerY] == 3){ playerX--; } break; case 40: // Down if (playerY < worldHeight) playerY++; if(board[playerX][playerY] == 3){ playerY--; } break; } viewX = playerX - Math.floor(0.5 * viewWidth); if (viewX < 0) viewX = 0; if (viewX+viewWidth > worldWidth) viewX = worldWidth - viewWidth; viewY = playerY - Math.floor(0.5 * viewHeight); if (viewY < 0) viewY = 0; if (viewY+viewHeight > worldHeight) viewY = worldHeight - viewHeight; }, false); My question is, is there a more efficient way of handling collisions, then loads of if statements for each key? The reason i ask is because i plan on having many items that the player will need to be able to pickup or not walk through like walls cliffs etc. Thanks for your time and help Tom

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• How to execute a Ruby file in Java, capable of calling functions from the Java program and receiving primitive-type results?

  - by Omega

  I do not fully understand what am I asking (lol!), well, in the sense of if it is even possible, that is. If it isn't, sorry. Suppose I have a Java program. It has a Main and a JavaCalculator class. JavaCalculator has some basic functions like public int sum(int a,int b) { return a + b } Now suppose I have a ruby file. Called MyProgram.rb. MyProgram.rb may contain anything you could expect from a ruby program. Let us assume it contains the following: class RubyMain def initialize print "The sum of 5 with 3 is #{sum(5,3)}" end def sum(a,b) # <---------- Something will happen here end end rubyMain = RubyMain.new Good. Now then, you might already suspect what I want to do: I want to run my Java program I want it to execute the Ruby file MyProgram.rb When the Ruby program executes, it will create an instance of JavaCalculator, execute the sum function it has, get the value, and then print it. The ruby file has been executed successfully. The Java program closes. Note: The "create an instance of JavaCalculator" is not entirely necessary. I would be satisfied with just running a sum function from, say, the Main class. My question: is such possible? Can I run a Java program which internally executes a Ruby file which is capable of commanding the Java program to do certain things and get results? In the above example, the Ruby file asks the Java program to do a sum for it and give the result. This may sound ridiculous. I am new in this kind of thing (if it is possible, that is). WHY AM I ASKING THIS? I have a Java program, which is some kind of game engine. However, my target audience is a bunch of Ruby coders. I don't want to have them learn Java at all. So I figured that perhaps the Java program could simply offer the functionality (capacity to create windows, display sprites, play sounds...) and then, my audience can simply code with Ruby the logic, which basically justs asks my Java engine to do things like displaying sprites or playing sounds. That's when I though about asking this.

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• Visual Studio 2010 Launch Events

  - by Jim Duffy

  Don’t miss out on the opportunity to learn about the new features in Visual Studio 2010. Check out the MSDN Events page and find out when the talented folks of the Developer & Evangelism group will be visiting your city to prove to you that /*Life Runs On Code*/. I’ll be attending the Raleigh event June 2, 2010 from 1:00 - 5:00 PM. North Carolina State University, Jane S. McKimmon Conference Center 1101 Gorman St Raleigh North Carolina 27606 United States From the Raleigh Event page: Event Overview Learn about the rich application platforms that Microsoft® Visual Studio® 2010 supports, including Windows® 7, the Web, SharePoint®, Windows Azure™, SQL®, and Windows® Phone 7 Series. From tighter tester and dev collaboration to new ALM tools, there’s a lot that’s new. Here’s what you can expect: Windows Development with Visual Studio 2010 Visual Studio has always been the best way to build compelling visual solutions for Windows. Visual Studio 2010 continues this trend with great new tooling support for Silverlight 4, WPF, and native development. In this demo heavy session, you’ll see how you can build rich Windows applications with Silverlight 4 using new trusted application features including out-of-browser execution, saving to the file system, and even COM Automation. You’ll also see how you can use the new Task Parallel Library from within a WPF application to take advantage of all those cores in today’s modern computers. Web and Cloud Development with Visual Studio 2010 If you build solutions for the web, then this session is for you. Come see how your existing skills move forward with Visual Studio 2010 both for in-house ASP.NET development and the new frontier of the Cloud. In this session, you’ll see improved designers, new HTML and JavaScript snippets, Web Forms enhancements, and how you can quickly build great web sites using Dynamic Data. You’ll see the changes made to testable web sites with MVC 2.0 and how we’ve integrated JQuery support into the platform. You’ll then see how easy it is to leverage your existing code and move to the cloud with Windows Azure. Windows Phone 7 Developer Tools and Platform Overview This session provides an overview of Visual Studio® 2010 for Windows Phone. Learn about the powerful capabilities of this new application platform and the developer tools experience including basic IDE usage, debugging, packaging, and deployment. This session also shows how you can use Microsoft Expression® Blend™ for Windows Phone to build great Silverlight applications. Have a day. :-|

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• ArchBeat Facebook Friday: Top 10 Shared Links - May 23-29, 2014

  - by OTN ArchBeat

  Among the 5,144 fans of the OTN ArchBeat Facebook Page the following Top 10 items were the most popular over the last seven days, May 23-29, 2014. GlassFish/Java EE Community Open Forum Today! | Reza Rahman Have questions about Glassfish? Java EE/GlassFish evangelist Reza Rahman has answers, and you can pick his brain tomorrow during an online forum organized by the London Glassfish User Group and C2B2. The event is free, but you must register in order to participate. Click the link for more information. Twitter Tuesday - Top 10 @ArchBeat Tweets - May 20-26, 2014 The top 10 @OTNArchBeat tweets for the week of May 20-26, 2014. Topics covered include ADF, Cloud, GoldenGate, KScope14, OBIEE, ODI, WebLogic, WebCenter, and more. FrameworkFolders Support has come to Oracle WebCenter Portal | JayJay Zheng Interested in working with Framework Folders in Oracle WebCenter Portal? Oracle ACE JayJay Zheng reviews the essentials. Video: Programming Best Practices - ADF Business Components | Frank Nimphius Frank Nimphius discusses best practices and recommendations for ADF Business Components in the latest video from ADF Architecture TV. Video: Kscope 2014 Preview: Data Modeling and Moving Meditation with Kent Graziano For your mind and your body! Oracle ACE Director Kent Graziano previews his Kscope 2014 data modeling presentations and the early morning Chi Gung sessions he will once again lead for Kscope attendees. OAG and OES Integration for Web API Security: skin and guts | Andre Correa A-Team architect Andre Correa's post examines a strategy for web API security that uses OAG (Oracle API Gateway) and OES (Oracle Entitlements Server). Getting Started with Coherence*Web in WebLogic Server 12.1.2 | Tim Middleton Solution architect Tim Middleton shows you how to configure Coherence*Web in WebLogic Server 12.1.2 and deploy a basic web application. SOA and Business Processes: You are the Process! Part of the 13-part "Industrial SOA" article series, this article looks at best practices for modeling and managing effective business processes. Authentication in Oracle Identity Federation/ IdP | Damien Carru Damien Carru discuss authentication when OIF acts as an IdP and how the server can be configured to use specific OAM Authentication Schemes to challenge the user. Caveats on Using WebLogic Server with JDK7 | JayJay Zheng Quick tech tips from Oracle ACE JayJay Zheng.

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• Have you used the ExecutionValue and ExecValueVariable properties?

  The ExecutionValue execution value property and it’s friend ExecValueVariable are a much undervalued feature of SSIS, and many people I talk to are not even aware of their existence, so I thought I’d try and raise their profile a bit. The ExecutionValue property is defined on the base object Task, so all tasks have it available, but it is up to the task developer to do something useful with it. The basic idea behind it is that it allows the task to return something useful and interesting about what it has performed, in addition to the standard success or failure result. The best example perhaps is the Execute SQL Task which uses the ExecutionValue property to return the number of rows affected by the SQL statement(s). This is a very useful feature, something people often want to capture into a variable, and start using the result set options to do. Unfortunately we cannot read the value of a task property at runtime from within a SSIS package, so the ExecutionValue property on its own is a bit of a let down, but enter the ExecValueVariable and we have the perfect marriage. The ExecValueVariable is another property exposed through the task (TaskHost), which lets us select a SSIS package variable. What happens now is that when the task sets the ExecutionValue, the interesting value is copied into the variable we set on the ExecValueVariable property, and a variable is something we can access and do something with. So put simply if the ExecutionValue property value is of interest, make sure you create yourself a package variable and set the name as the ExecValueVariable. Have  look at the 3 step guide below: 1 Configure your task as normal, for example the Execute SQL Task, which here calls a stored procedure to do some updates. 2 Create variable of a suitable type to match the ExecutionValue, an integer is used to match the result we want to capture, the number of rows. 3 Set the ExecValueVariable for the task, just select the variable we created in step 2. You need to do this in Properties grid for the task (Short-cut key, select the task and press F4) Now when we execute the sample task above, our variable UpdateQueueRowCount will get the number of rows we updated in our Execute SQL Task. I’ve tried to collate a list of tasks that return something useful via the ExecutionValue and ExecValueVariable mechanism, but the documentation isn’t always great. Task ExecutionValue Description Execute SQL Task Returns the number of rows affected by the SQL statement or statements. File System Task Returns the number of successful operations performed by the task. File Watcher Task Returns the full path of the file found Transfer Error Messages Task Returns the number of error messages that have been transferred Transfer Jobs Task Returns the number of jobs that are transferred Transfer Logins Task Returns the number of logins transferred Transfer Master Stored Procedures Task Returns the number of stored procedures transferred Transfer SQL Server Objects Task Returns the number of objects transferred WMI Data Reader Task Returns an object that contains the results of the task. Not exactly clear, but I assume it depends on the WMI query used.

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• As a tooling/automation developer, can I be making better use of OOP?

  - by Tom Pickles

  My time as a developer (~8 yrs) has been spent creating tooling/automation of one sort or another. The tools I develop usually interface with one or more API's. These API's could be win32, WMI, VMWare, a help-desk application, LDAP, you get the picture. The apps I develop could be just to pull back data and store/report. It could be to provision groups of VM's to create live like mock environments, update a trouble ticket etc. I've been developing in .Net and I'm currently reading into design patterns and trying to think about how I can improve my skills to make better use of and increase my understanding of OOP. For example, I've never used an interface of my own making in anger (which is probably not a good thing), because I honestly cannot identify where using one would benefit later on when modifying my code. My classes are usually very specific and I don't create similar classes with similar properties/methods which could use a common interface (like perhaps a car dealership or shop application might). I generally use an n-tier approach to my apps, having a presentation layer, a business logic/manager layer which interfaces with layer(s) that make calls to the API's I'm working with. My business entities are always just method-less container objects, which I populate with data and pass back and forth between my API interfacing layer using static methods to proxy/validate between the front and the back end. My code by nature of my work, has few common components, at least from what I can see. So I'm struggling to see how I can better make use of OOP design and perhaps reusable patterns. Am I right to be concerned that I could be being smarter about how I work, or is what I'm doing now right for my line of work? Or, am I missing something fundamental in OOP? EDIT: Here is some basic code to show how my mgr and api facing layers work. I use static classes as they do not persist any data, only facilitate moving it between layers. public static class MgrClass { public static bool PowerOnVM(string VMName) { // Perform logic to validate or apply biz logic // call APIClass to do the work return APIClass.PowerOnVM(VMName); } } public static class APIClass { public static bool PowerOnVM(string VMName) { // Calls to 3rd party API to power on a virtual machine // returns true or false if was successful for example } }

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• Why would you dual-run an app on Azure and AWS?

  - by Elton Stoneman

  Originally posted on: http://geekswithblogs.net/EltonStoneman/archive/2013/11/10/why-would-you-dual-run-an-app-on-azure-and-aws.aspxI had this question from a viewer of my Pluralsight course, Implementing the Reactive Manifesto with Azure and AWS, and thought I’d publish the response. So why would you dual-run your cloud app by hosting it on Azure and AWS? Sounds like a lot of extra development and management overhead. Well the most compelling reasons are reliability and portability. In 2012 I was working for a client who was making a big investment in the cloud, and at the end of the year we published their first external API for business partners. It was hosted in Azure and used some really nice features to route back into existing on-premise services. We were able to publish a clean, simple API to partners, and hide away the underlying complexity of the internal services while still leveraging them to do all the work. Two days after we went live, we were hit by the Azure SSL certificate expiry outage, and our API was unavailable for the best part of 3 days. Fortunately we had planned a gradual roll-out to partners, so the impact was minimal, but we’d been intending to ramp up quickly, and if the outage had happened a week or two later we would have been in a very bad place. Not least because our app could only run on Azure, we couldn’t package it up for another service without going back and reworking the code. More recently AWS had an issue with a networking device in one of their data centres which caused an outage that took the best part of a day to resolve. In both scenarios the SLAs are worthless, as you’ll get back a small percentage of your cloud expenditure, which is going to be negligible compared to your costs in dealing with the outage. And if your app is built specifically for AWS or Azure then if there’s an extended outage you can’t just deploy it onto a new set of kit from a different supplier. And the chances are pretty good there will be another extended outage, both for Microsoft and for Amazon. But the chances are small that it will happen to both at the same time. So my basic guidance has been: ignore the SLAs, go for better uptime by using two clouds. As soon as you need to scale beyond a single instance, start by scaling out to another cloud. Then scale out to different data centres in both clouds. Then you’ve got dual-cloud, quadruple-datacentre redundancy, so any more scaling you need can be left to the clouds to auto-scale themselves. By running in both clouds, you’ve made your app portable, so in the highly unlikely event that both AWS and Azure go down in multiple regions, you’ll have a deployment package which will let you spin up a new stack on yet another cloud, without having to rework your solution.

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• Writing or extending existing emacs packages: is it worth or should I move to Netbeans/Eclipse?

  - by Andrea

  I'm finishing my master degree course in CS and I've almost become addicted to Emacs. I've used it to write in C, Latex, Java, JSP,XML, CommonLisp, Ada and other languages no other editor supported, like AMPL. I'd like to improve the packages I've been using the most or create new ones, but, in practice, I find that the implementation of Emacs leaves a lot to be desired. There are a lot of poorly-featured/poorly-maintained packages with either overlapping functionalities or obscure incompatibilities, and Elisp just seems to foster the situation by lacking the common features modern lisps have. In contrast Eclipse and Netbeans are actively improved and it does seem they can be effective for non-mainstream languages. I tried Hibachi for Ada in Eclipse and it worked well, there's CUPS for Lisp in Eclipse and LambdaBeans built using NetBeans components. On the other hand those plugins seem to be less active than their Emacs' counterparts, for example Hibachi was archived last year. What's your opinion on this? Which editor should I write extension for? EDIT: To answer Larry Coleman (see comment below): I like Emacs as a user because it is efficient both for me and the computer I'm using. It's fast and the textual interface (i.e. minibuffer) allows for quick interaction. It's solid and packages are usually small and easy to manage. If I need to correct or remove something I usually just have to change a row in my .emacs or an elisp file, or delete a directory. Eclipse plugins rely on a more complicated process that screwed my Eclipse configuration a couple of times, forcing me to do a clean reinstall. Emacs works as long as I use the basic packages. If I need something more complicated the situation gets pretty hairy. As a "power user" I think that the best I can hope for is to write a severely crippled version of the extensions I'd actually like to have; in other words, that it's not worth the trouble. I'd like to write extensions for the things I'd like to have automated in Emacs, for example project support with automated tag-table update on file writing. There are a few projects on this that lack integration, documentation, extensibility and so forth. The best one is probably CEDET, for which I believe the Greenspun's 10th rule can be applied. EDIT: To comment Larry Coleman's answer I'm pretty sure I can pick elisp programming but the extensions I have in mind don't exist yet despite their relative simplicity and the effort more knowledgeable people poured into related projects.This makes me wonder whether it is so because of the way emacs is developed, i.e. people tend to write their own little extensions without coordination, or its implementation, its extension language not being able to keep up with the growing complexity.

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• EBS – ATG Webcast 9/11 - 9/12

  - by cwarticki

  EBS – ATG Webcast in September 2012 EBS – Multiple Language Support (MLS) Agenda :EBS is MLS Ready                                                                                 NLS / MLS Basic ArchitectureNLS / MLS InstallationNLS / MLS Configuration Settings                                                                    TroubleshootingQuestion and AnswersEMEA Session : September 11, 2012 at 09:00 UK / 10:00 CET / 13:30 India / 17:00 Japan / 18:00 Sydney (Australia) Details & Registration : Note 1480084.1 Direct link to register in WebEx US Session : September 12, 2012 at 18:00 UK / 19:00 CET / 10:00 AM Pacific / 11:00 AM Mountain/ 01:00 PM Eastern ·      Details & Registration : Note 1480085.1 ·      Direct link to register in WebEx ·         Schedules, recordings and the Presentations of the Advisor Webcast drove under the EBS Applications Technology area can be found in Note 1186338.1. ·         Current Schedules of Advisor Webcast for all Oracle Products can be found on Note 740966.1 ·         Post Presentation Recordings of the Advisor Webcasts for all Oracle Products can be found on Note 740964.1 If you have any question about the schedules or if you have a suggestion for an Advisor Webcast to be planned in future, please send an E-Mail to Ruediger Ziegler.

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• Is it ok to initialize an RB_ConstraintActor in PostBeginPlay?

  - by Almo

  I have a KActorSpawnable subclass that acts weird. In PostBeginPlay, I initialize an RB_ConstraintActor; the default is not to allow rotation. If I create one in the editor, it's fine, and won't rotate. If I spawn one, it rotates. Here's the class: class QuadForceKActor extends KActorSpawnable placeable; var(Behavior) bool bConstrainRotation; var(Behavior) bool bConstrainX; var(Behavior) bool bConstrainY; var(Behavior) bool bConstrainZ; var RB_ConstraintActor PhysicsConstraintActor; simulated event PostBeginPlay() { Super.PostBeginPlay(); PhysicsConstraintActor = Spawn(class'RB_ConstraintActorSpawnable', self, '', Location, rot(0, 0, 0)); if(bConstrainRotation) { PhysicsConstraintActor.ConstraintSetup.bSwingLimited = true; PhysicsConstraintActor.ConstraintSetup.bTwistLimited = true; } SetLinearConstraints(bConstrainX, bConstrainY, bConstrainZ); PhysicsConstraintActor.InitConstraint(self, None); } function SetLinearConstraints(bool InConstrainX, bool InConstrainY, bool InConstrainZ) { if(InConstrainX) { PhysicsConstraintActor.ConstraintSetup.LinearXSetup.bLimited = 1; } else { PhysicsConstraintActor.ConstraintSetup.LinearXSetup.bLimited = 0; } if(InConstrainY) { PhysicsConstraintActor.ConstraintSetup.LinearYSetup.bLimited = 1; } else { PhysicsConstraintActor.ConstraintSetup.LinearYSetup.bLimited = 0; } if(InConstrainZ) { PhysicsConstraintActor.ConstraintSetup.LinearZSetup.bLimited = 1; } else { PhysicsConstraintActor.ConstraintSetup.LinearZSetup.bLimited = 0; } } DefaultProperties { bConstrainRotation=true bConstrainX=false bConstrainY=false bConstrainZ=false bSafeBaseIfAsleep=false bNoEncroachCheck=false } Here's the code I use to spawn one. It's a subclass of the one above, but it doesn't reference the constraint at all. local QuadForceKCreateBlock BlockActor; BlockActor = spawn(class'QuadForceKCreateBlock', none, 'PowerCreate_Block', BlockLocation(), m_PreparedRotation, , false); BlockActor.SetDuration(m_BlockDuration); BlockActor.StaticMeshComponent.SetNotifyRigidBodyCollision(true); BlockActor.StaticMeshComponent.ScriptRigidBodyCollisionThreshold = 0.001; BlockActor.StaticMeshComponent.SetStaticMesh(m_ValidCreationBlock.StaticMesh); BlockActor.StaticMeshComponent.AddImpulse(m_InitialVelocity); I used to initialize an RB_ConstraintActor where I spawned it from the outside. This worked, which is why I'm pretty sure it has nothing to do with the other code in QuadForceKCreateBlock. I then added the internal constraint in QuadForceKActor for other purposes. When I realized I had two constraints on the CreateBlock doing the same thing, I removed the constraint code from the place where I spawn it. Then it started rotating. Is there a reason I should not be initializing an RB_ConstraintActor in PostBeginPlay? I feel like there's some basic thing about how the engine works that I'm missing.

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• Confusion about inheritance

  - by Samuel Adam

  I know I might get downvoted for this, but I'm really curious. I was taught that inheritance is a very powerful polymorphism tool, but I can't seem to use it well in real cases. So far, I can only use inheritance when the base class is an abstract class. Examples : If we're talking about Product and Inventory, I quickly assumed that a Product is an Inventory because a Product must be inventorized as well. But a problem occured when user wanted to sell their Inventory item. It just doesn't seem to be right to change an Inventory object to it's subtype (Product), it's almost like trying to convert a parent to it's child. Another case is Customer and Member. It is logical (at least for me) to think that a Member is a Customer with some more privileges. Same problem occurred when user wanted to upgrade an existing Customer to become a Member. A very trivial case is the Employee case. Where Manager, Clerk, etc can be derived from Employee. Still, the same upgrading issue. I tried to use composition instead for some cases, but I really wanted to know if I'm missing something for inheritance solution here. My composition solution for those cases : Create a reference of Inventory inside a Product. Here I'm making an assumption about that Product and Inventory is talking in a different context. While Product is in the context of sales (price, volume, discount, etc), Inventory is in the context of physical management (stock, movement, etc). Make a reference of Membership instead inside Customer class instead of previous inheritance solution. Therefor upgrading a Customer is only about instantiating the Customer's Membership property. This example is keep being taught in basic programming classes, but I think it's more proper to have those Manager, Clerk, etc derived from an abstract Role class and make it a property in Employee. I found it difficult to find an example of a concrete class deriving from another concrete class. Is there any inheritance solution in which I can solve those cases? Being new in this OOP thing, I really really need a guidance. Thanks!

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• Pro SharePoint 2010 Business Intelligence Solutions

  - by Sahil Malik

  Ad:: SharePoint 2007 Training in .NET 3.5 technologies (more information). Oh yeah baby, it’s out finally! This book is what I wanted to write for so long now, but never really got a chance to. For SharePoint 2007, I authored the SharePoint section of “Smart BI Solutions with SQL Server 2008” for MS Press. But never really got the time, to author a full book that this topic deserved. Until SharePoint 2010, we actually have a full book on this topic. So first things first, I didn’t actually write it. My role was limited to the overall concept, the outline, the layout, completion of it, code samples, identifying what we need in here, vouching for technical accuracy, identifying authors etc. The real work was done by Srini (5 chapters), and Steve (1 chapter). So credit given where it is due. But, with that said, this is a pretty good book. It has always been a challenge to find the superman that knows both, data ware housing concepts, and SharePoint concepts. The data ware housing concepts include basic stuff you need to know to work in the BI area such as cubes, MDX queries, etc. So chapter 1 covers that – and if you’re a hardcore DBA, feel free to skip Chapter 1. Then beyond that, we take every single SharePoint 2010 BI topic, and slice and dice it in detail. The topics we deal with are - Visio Services Reporting services Business Connectivity Services Excel Services PerformancePoint Services And in covering each of these topics, we ensure that a general layout was followed for each topic, to ensure completeness of content. We make sure we cover Setup related issues and advice Point and click usage Code usage, i.e. extensibility using visual studio and a walkthrough of the administration side of things, including powershell. (Yes, I insisted on that in being there in every chapter). Writing a book is always a lot of work, so we hope you find it useful. And it should go very well with the other book I just reviewed, which is Microsoft ADO.NET 4, step by step. Comment on the article ....

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• WIF, ADFS 2 and WCF&ndash;Part 1: Overview

  - by Your DisplayName here!

  A lot has been written already about passive federation and integration of WIF and ADFS 2 into web apps. The whole active/WS-Trust feature area is much less documented or covered in articles and blogs. Over the next few posts I will try to compile all relevant information about the above topics – but let’s start with an overview. ADFS 2 has a number of endpoints under the /services/trust base address that implement the WS-Trust protocol. They are grouped by the WS-Trust version they support (/13 and /2005), the client credential type (/windows*, /username*, /certificate*) and the security mode (*transport, *mixed and message). You can see the endpoints in the MMC console under the Service/Endpoints page. So in other words, you use one of these endpoints (which exactly depends on your configuration / system setup) to request tokens from ADFS 2. The bindings behind the endpoints are more or less standard WCF bindings, but with SecureConversation (establishSecurityContext) disabled. That means that whenever you need to programmatically talk to these endpoints – you can (easily) create client bindings that are compatible. Another option is to use the special bindings that come with WIF (in the Microsoft.IdentityModel.Protocols.WSTrust.Bindings namespace). They are already pre-configured to be compatible with the ADFS endpoints. The downside of these bindings is, that you can’t use them in configuration. That’s definitely a feature request of mine for the next version of WIF. The next important piece of information is the so called Federation Service Identifier. This is the value that you (at least by default) have to use as a realm/appliesTo whenever you are requesting a token for ADFS (e.g. in  IdP –> RSTS scenario). Or (even more) technically speaking, ADFS 2 checks for this value in the audience URI restriction in SAML tokens. You can get to this value by clicking the “Edit Federation Service Properties” in the MMC when the Service tree-node is selected. OK – I will come back to this basic information in the following posts. Basically I want to go through the following scenarios: ADFS in the IdP role ADFS in the R-STS role (with a chained claims provider) Using the WCF bindings for automatic token issuance Using WSTrustChannelFactory for manual token handling Stay tuned…

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• design a model for a system of dependent variables

  - by dbaseman

  I'm dealing with a modeling system (financial) that has dozens of variables. Some of the variables are independent, and function as inputs to the system; most of them are calculated from other variables (independent and calculated) in the system. What I'm looking for is a clean, elegant way to: define the function of each dependent variable in the system trigger a re-calculation, whenever a variable changes, of the variables that depend on it A naive way to do this would be to write a single class that implements INotifyPropertyChanged, and uses a massive case statement that lists out all the variable names x1, x2, ... xn on which others depend, and, whenever a variable xi changes, triggers a recalculation of each of that variable's dependencies. I feel that this naive approach is flawed, and that there must be a cleaner way. I started down the path of defining a CalculationManager<TModel> class, which would be used (in a simple example) something like as follows: public class Model : INotifyPropertyChanged { private CalculationManager<Model> _calculationManager = new CalculationManager<Model>(); // each setter triggers a "PropertyChanged" event public double? Height { get; set; } public double? Weight { get; set; } public double? BMI { get; set; } public Model() { _calculationManager.DefineDependency<double?>( forProperty: model => model.BMI, usingCalculation: (height, weight) => weight / Math.Pow(height, 2), withInputs: model => model.Height, model.Weight); } // INotifyPropertyChanged implementation here } I won't reproduce CalculationManager<TModel> here, but the basic idea is that it sets up a dependency map, listens for PropertyChanged events, and updates dependent properties as needed. I still feel that I'm missing something major here, and that this isn't the right approach: the (mis)use of INotifyPropertyChanged seems to me like a code smell the withInputs parameter is defined as params Expression<Func<TModel, T>>[] args, which means that the argument list of usingCalculation is not checked at compile time the argument list (weight, height) is redundantly defined in both usingCalculation and withInputs I am sure that this kind of system of dependent variables must be common in computational mathematics, physics, finance, and other fields. Does someone know of an established set of ideas that deal with what I'm grasping at here? Would this be a suitable application for a functional language like F#? Edit More context: The model currently exists in an Excel spreadsheet, and is being migrated to a C# application. It is run on-demand, and the variables can be modified by the user from the application's UI. Its purpose is to retrieve variables that the business is interested in, given current inputs from the markets, and model parameters set by the business.

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• What could be the best way to generalize data from Facebook and Twitter?

  - by Sjaak van der Heide

  I am not sure if this is the best subsite to ask this question, but I'm pretty sure it doesn't fit on the normal or facebook SO page... I've been asked to make a general API for connecting to several Social Media platforms (at the moment Facebook and Twitter). I have already realised both of them seperately. Meaning I retrieve the data I need from both Facebook and Twitter and hold the data in it's own dataclass. In my case a list of FacebookTimelineItems and a list of TwitterTimelineItems. now the hard part is taking the parts that are used in both (username, id, message and such) and make 1 general class that is eventually passed on to who/whatever sent the call to my API. these are two pics of the data classes I have: http://imageshack.us/photo/my-images/703/facebookdata.png/ http://imageshack.us/photo/my-images/204/twitterdata.png/ probably not 100% correct but it gives an idea what it looks like. Now I've been having several idea about how to go about and generalize the two, which is harder then I thought at first. Create an interface (TimelineItem) and let the other classes extend that one. this way I'll always be sure I have a class that contains at least the basic info I need. downside is that deserializing the JSON seems to be a nightmare. Use the two dataclasses I have and combine them into a new class afterwards, then pass that one back to whoever requested it. This would probably work but I get the idea it's not the best way to tackle this problem, and is pretty dodgy IF I get it working. Or, in case of the other two being nearly impossible. Keep the two seperated in the front end, and go sit in the corner crying because I've just figured out you can't lump together facebook and twitter... Note: I don't have to make the front end part (view), I just make sure the Model is nicely filled with data :) I hope I placed this in the right section, if I didn't I apologise and would like to know where I should go with my question. Thanks in advance for any replied/ideas/opinions on this.

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• Certification Notes: 70-583 Designing and Developing Windows Azure Applications

  - by BuckWoody

  Last Updated: 02/01/2011 It’s time for another certification, and we’ve just release the 70-583 exam on Windows Azure. I’ve blogged my “study plans” here before on other certifications, so I thought I would do the same for this one. I’ll also need to take exam 70-513 and 70-516; but I’ll post my notes on those separately. None of these are “brain dumps” or any questions from the actual tests - just the books, links and notes I have from my studies. I’ll update these references as I’m studying, so bookmark this site and watch my Twitter and Facebook posts for when I’ll update them, or just subscribe to the RSS feed. A “Green” color on the check-block means I’ve done that part so far, red means I haven’t. First, I need to refresh my memory on some basic coding, so along with the Azure-specific information I’m reading the following general programming books: Introducing Microsoft .NET (Pro-Developer): link   Head First C#, 2E: A Learner's Guide to Real-World Programming with Visual C# and .NET: link Microsoft Visual C# 2008 Step by Step: link  c The first place to start is at the official site for the certification. link c On that page you’ll find several resources, and the first you should follow is the “Save to my learning” so you have a place to track everything. Then click the “Related Learning Plans” link and follow the videos and read the documentation in each of those bullets. There are six areas on the learning plan that you should focus on - make sure you open the learning plan to drill into the specifics. c Designing Data Storage Architecture (18%) Books I’m Reading: Links: My Notes: c Optimizing Data Access and Messaging (17%) Books I’m Reading: Links: My Notes: c Designing the Application Architecture (19%) Books I’m Reading: Applied Architecture Patterns on the Microsoft Platform: link Links: My Notes: c Preparing for Application and Service Deployment (15%) Books I’m Reading: Links: My Notes: c Investigating and Analyzing Applications (16%) Books I’m Reading: Links: My Notes: c Designing Integrated Solutions (15%) Books I’m Reading: Applied Architecture Patterns on the Microsoft Platform (2nd mention) Links: My Notes:

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• Support ARMv7 instruction set in Windows Embedded Compact applications

  - by Valter Minute

  On of the most interesting new features of Windows Embedded Compact 7 is support for the ARMv5, ARMv6 and ARMv7 instruction sets instead of the ARMv4 “generic” support provided by the previous releases. This means that code build for Windows Embedded Compact 7 can leverage features (like the FPU unit for ARMv6 and v7) and instructions of the recent ARM cores and improve their performances. Those improvements are noticeable in graphics, floating point calculation and data processing. The ARMv7 instruction set is supported by the latest Cortex-A8, A9 and A15 processor families. Those processor are currently used in tablets, smartphones, in-car navigation systems and provide a great amount of processing power and a low amount of electric power making them very interesting for portable device but also for any kind of device that requires a rich user interface, processing power, connectivity and has to keep its power consumption low. The bad news is that the compiler provided with Visual Studio 2008 does not provide support for ARMv7, building native applications using just the ARMv4 instruction set. Porting a Visual Studio “Smart Device” native C/C++ project to Platform Builder is not easy and you’ll lack many of the features that the VS2008 application development environment provides. You’ll also need access to the BSP and OSDesign configuration for your device to be able to build and debug your application inside Platform Builder and this may prevent independent software vendors from using the new compiler to improve their applications performances. Adeneo Embedded now provides a whitepaper and a Visual Studio plug-in that allows usage of the new ARMv7 enabled compiler to build applications inside Visual Studio 2008. I worked on the whitepaper and the tools, with the help of my colleagues and now the results can be downloaded from Adeneo Embedded’s website: http://www.adeneo-embedded.com/OS-Technologies/Windows-Embedded (Click on the “WEC7 ARMv7 Whitepaper tab to access the download links, free registration required) A very basic benchmark showed a very good performance improvement in integer and floating-point operations. Obviously your mileage may vary and we can’t promise the same amount of improvement on any application, but with a small effort on your side (even smaller if you use the plug-in) you can try on your own application. ARMv7 support is provided using Platform Builder’s compiler and VS2008 application debugger is not able to debut ARMv7 code, so you may need to put in place some workaround like keeping ARMv4 code for debugging etc.

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