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• Jitter during wall collisions with Bullet Physics: contact/penetration tolerance?

  - by Niriel

  I use the bullet physics engine through Panda3d. My scene is still very simple, think 'Wolfenstein3d': tile-based, walls are solid cubes. I expect walls to block the player, and I expect the player to slide along the walls in case of non-normal incidence. What I get is what I expect, with one difference: there is some jitter. If I try to force myself into the wall, then I see the frames blinking quickly between two positions. These differ by about 0.04 units of distance, which corresponds to 4 cm in my game. I noticed a 4 cm elsewhere: the bottom of my player capsule is 4 cm below ground, when at rest. Does that mean that there is somewhere in the Bullet engine a default 0.04-units-long tolerance to differentiate contact from collision? If so, what should I do ? Should I change the scale of my game so that these 0.04 units correspond to 0.4 cm, making the jitter ten times smaller? Or can I ask bullet to change its tolerance to a smaller value? Edit This is the jitter I get: 6.155 - 6.118 = 0.036 LPoint3f(0, 6.11694, 0.835) LPoint3f(0, 6.15499, 0.835) LPoint3f(0, 6.11802, 0.835) LPoint3f(0, 6.15545, 0.835) LPoint3f(0, 6.11817, 0.835) LPoint3f(0, 6.15726, 0.835) LPoint3f(0, 6.11876, 0.835) LPoint3f(0, 6.15911, 0.835) LPoint3f(0, 6.11937, 0.835) I found a setMargin method. I set it to 5 mm both on the BoxShape for the walls and on the Capsule shape for the player. It still jitters by about 35 mm as illustrated by this log (11.117 - 11.082 = 0.035): LPoint3f(0, 11.0821, 0.905) LPoint3f(0, 11.1169, 0.905) LPoint3f(0, 11.082, 0.905) LPoint3f(0, 11.117, 0.905) LPoint3f(0, 11.082, 0.905) LPoint3f(0, 11.117, 0.905) LPoint3f(0, 11.0821, 0.905) LPoint3f(0, 11.1175, 0.905) LPoint3f(0, 11.0822, 0.905) LPoint3f(0, 11.1178, 0.905) LPoint3f(0, 11.0823, 0.905) LPoint3f(0, 11.1183, 0.905) The margin on the capsule did change my penetration with the floor though, I'm a bit higher (0.905 instead of 0.835). However, it did not change anything when colliding with the walls. How can I make the collisions against the walls less jittery? Edit, the day after: After more investigation, it appears that dynamic objects behave well. My problem comes from the btKinematicCharacterController that I use for moving my character; that stuff is totally bugged, according to the whole Internet :/.

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• Profiling Startup Of VS2012 &ndash; Ants Profiler

  - by Alois Kraus

  I just downloaded ANTS Profiler 7.4 to check how fast it is and how deep I can analyze the startup of Visual Studio 2012. The Pro version which is useful does cost 445€ which is ok. To measure a complex system I decided to simply profile VS2012 (Update 1) on my older Intel 6600 2,4GHz with 3 GB RAM and a 32 bit Windows 7. Ants Profiler is really easy to use. So lets try it out. The Ants Profiler does want to start the profiled application by its own which seems to be rather common. I did choose Method Level timing of all managed methods. In the configuration menu I did want to get all call stacks to get full details. Once this is configured you are ready to go.   After that you can select the Method Grid to view Wall Clock Time in ms. I hate percentages which are on by default because I do want to look where absolute time is spent and not something else.   From the Method Grid I can drill down to see where time is spent in a nice and I can look at the decompiled methods where the time is spent. This does really look nice. But did you see the size of the scroll bar in the method grid? Although I wanted all call stacks I do get only about 4 pages of methods to drill down. From the scroll bar count I would guess that the profiler does show me about 150 methods for the complete VS startup. This is nonsense. I will never find a bottleneck in VS when I am presented only a fraction of the methods that were actually executed. I have also tried in the configuration window to also profile the extremely trivial functions but there was no noticeable difference. It seems that the Ants Profiler does filter away way too many details to be useful for bigger systems. If you want to optimize a CPU bound operation inside NUnit then Ants Profiler is with its line level timings a very nice tool to work with. But for bigger stuff it is certainly not usable. I also do not like that I must start the profiled application from the profiler UI. This makes it hard to profile processes which are started by some other process. Next: JetBrains dotTrace

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• A little gem from MPN&ndash;FREE online course on Architectural Guidance for Migrating Applications to Windows Azure Platform

  - by Eric Nelson

  I know a lot of technical people who work in partners (ISVs, System Integrators etc). I know that virtually none of them would think of going to the Microsoft Partner Network (MPN) learning portal to find some deep and high quality technical content. Instead they would head to MSDN, Channel 9, msdev.com etc. I am one of those people :-) Hence imagine my surprise when i stumbled upon this little gem Architectural Guidance for Migrating Applications to Windows Azure Platform (your company and hence your live id need to be a member of MPN – which is free to join). This is first class stuff – and represents about 4 hours which is really 8 if you stop and ponder :) Course Structure The course is divided into eight modules.  Each module explores a different factor that needs to be considered as part of the migration process. Module 1:  Introduction:  This section provides an introduction to the training course, highlighting the values of the Windows Azure Platform for developers. Module 2:  Dynamic Environment: This section goes into detail about the dynamic environment of the Windows Azure Platform. This session will explain the difference between current development states and the Windows Azure Platform environment, detail the functions of roles, and highlight development considerations to be aware of when working with the Windows Azure Platform. Module 3:  Local State: This session details the local state of the Windows Azure Platform. This section details the different types of storage within the Windows Azure Platform (Blobs, Tables, Queues, and SQL Azure). The training will provide technical guidance on local storage usage, how to write to blobs, how to effectively use table storage, and other authorization methods. Module 4:  Latency and Timeouts: This session goes into detail explaining the considerations surrounding latency, timeouts and how to assess an IT portfolio. Module 5:  Transactions and Bandwidth: This session details the performance metrics surrounding transactions and bandwidth in the Windows Azure Platform environment. This session will detail the transactions and bandwidth costs involved with the Windows Azure Platform and mitigation techniques that can be used to properly manage those costs. Module 6:  Authentication and Authorization: This session details authentication and authorization protocols within the Windows Azure Platform. This session will detail information around web methods of authorization, web identification, Access Control Benefits, and a walkthrough of the Windows Identify Foundation. Module 7:  Data Sensitivity: This session details data considerations that users and developers will experience when placing data into the cloud. This section of the training highlights these concerns, and details the strategies that developers can take to increase the security of their data in the cloud. Module 8:  Summary Provides an overall review of the course.

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• Ubuntu doesn't "see" external USB Hard Disk

  - by Mina Michael

  It's NTFS. It's USB2. I'm using Ubuntu 13.04. It works perfectly fine on Windows (which excludes cable and hardware problems). I have two Ubuntu computers and it's not detected on either. It's about 500 GB. Edits: Following the first link, I input sudo lsusb in a terminal; before and after connecting the HDD. The difference was Bus 001 Device 012: ID 14cd:6116 Super Top M6116 SATA Bridge. There it is! ("sata bridge" used to appear in a windows notification when I plugged in the HDD in!). ...This means that Ubuntu detects it but is it not mounting it? I tried this: sudo mount /dev/sdb1 /mnt But gives this: mount: special device /dev/sdb1 does not exist I also tried: sudo mount /dev/sdc1 /mnt but it stays with no output forever. I left it in background for about 30 min.s. sudo fdisk -l gives out this: Disk /dev/sda: 160.0 GB, 160041885696 bytes 255 heads, 63 sectors/track, 19457 cylinders, total 312581808 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: 0xa42d04a3 Device Boot Start End Blocks Id System /dev/sda1 63 80324 40131 de Dell Utility /dev/sda2 * 80325 102481919 51200797+ 7 HPFS/NTFS/exFAT /dev/sda3 263874558 312580095 24352769 5 Extended /dev/sda4 102481920 263872511 80695296 7 HPFS/NTFS/exFAT /dev/sda5 263874560 310505471 23315456 83 Linux /dev/sda6 310507520 312580095 1036288 82 Linux swap / Solaris Partition table entries are not in disk order Disk /dev/sdc: 500.1 GB, 500107862016 bytes 255 heads, 63 sectors/track, 60801 cylinders, total 976773168 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: 0x5822aaea Device Boot Start End Blocks Id System /dev/sdc1 2048 976769023 488383488 7 HPFS/NTFS/exFAT The part below "Partition table entries are not in disk order" takes about 5 minutes to appear. The outputs of ls /dev/ | grep sd before and after connecting the HDD: before: sda sda1 sda2 sda3 sda4 sda5 sda6 ,after: sda sda1 sda2 sda3 sda4 sda5 sda6 sdd sdd1 The second output has the lines sdd and sdd1 different from the first one. IT SHOWED THE FILES!! The command sudo mount /dev/sdd1 /mnt worked after I typed in sudo fdisk -l!!! Thanks a million!! :) :)

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• OpenGL ES 2 jittery camera movement

  - by user16547

  First of all, I am aware that there's no camera in OpenGL (ES 2), but from my understanding proper manipulation of the projection matrix can simulate the concept of a camera. What I'm trying to do is make my camera follow my character. My game is 2D, btw. I think the principle is the following (take Super Mario Bros or Doodle Jump as reference - actually I'm trying to replicate the mechanics of the latter): when the caracter goes beyond the center of the screen (in the positive axis/direction), update the camera to be centred on the character. Else keep the camera still. I did accomplish that, however the camera movement is noticeably jittery and I ran out of ideas how to make it smoother. First of all, my game loop (following this article): private int TICKS_PER_SECOND = 30; private int SKIP_TICKS = 1000 / TICKS_PER_SECOND; private int MAX_FRAMESKIP = 5; @Override public void run() { loops = 0; if(firstLoop) { nextGameTick = SystemClock.elapsedRealtime(); firstLoop = false; } while(SystemClock.elapsedRealtime() > nextGameTick && loops < MAX_FRAMESKIP) { step(); nextGameTick += SKIP_TICKS; loops++; } interpolation = ( SystemClock.elapsedRealtime() + SKIP_TICKS - nextGameTick ) / (float)SKIP_TICKS; draw(); } And the following code deals with moving the camera. I was unsure whether to place it in step() or draw(), but it doesn't make a difference to my problem at the moment, as I tried both and neither seemed to fix it. center just represents the y coordinate of the centre of the screen at any time. Initially it is 0. The camera object is my own custom "camera" which basically is a class that just manipulates the view and projection matrices. if(character.getVerticalSpeed() >= 0) { //only update camera if going up float[] projectionMatrix = camera.getProjectionMatrix(); if( character.getY() > center) { center += character.getVerticalSpeed(); cameraBottom = center + camera.getBottom(); cameraTop = center + camera.getTop(); Matrix.orthoM(projectionMatrix, 0, camera.getLeft(), camera.getRight(), center + camera.getBottom(), center + camera.getTop(), camera.getNear(), camera.getFar()); } } Any thought about what I should try or what I am doing wrong? Update 1: I think I updated every value you can see on screen to check whether the jittery movement is affected by that, but nothing changed, so something must be fundamentally flawed with my approach/calculations.

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• IPS Package Groups

  - by Alan_Solaris_RE

  IPS group packages consist solely of dependencies on other packages that make up a logical grouping of software. These are similar to, but not the equivalent of, Solaris 10 metaclusters. The main difference is that metaclusters are nested subsets ranging from a minimal install to nearly all packages on the media. Group packages have no such hierarchy. They can overlap other groups, or be completely disjoint sets. A group dependency is set this way in an IPS package manifest file: depend fmri=full/pkg/name type=group Current Solaris Groups Solaris currently has 4 system groups defined. These are used for different types of installation, and are included in the xml manifest files used by the various Solaris installers: Package Name Summary Description Default Installation For:  group/system/solaris-desktop Oracle Solaris Desktop Provides an Oracle Solaris desktop environment Live Media  group/system/solaris-large-server Oracle Solaris Large Server Provides an Oracle Solaris large server environment Text Installer  group/system/solaris-small-server Oracle Solaris Small Server Provides a useful command-line Oracle Solaris environment  Zones  group/system/solaris-auto-install  Oracle Solaris Automated Installer Client  Provides an Oracle Solaris Automated Installer client  Automated Installer There are also several "feature" groups such as AMP and GNU Developer Tools. These are provided for convenience, but are not used directly by any installers. Retrieving Group Package Information A listing of all current groups can be found with the command: pkg info -r group/* A listing of all the packages in a group can be obtained with: pkg contents -o fmri -H -rt depend -a type=group groupname An example: $ pkg contents -o fmri -H -rt depend -a type=group solaris-desktop archiver/gnu-tar audio/audio-utilities codec/flac codec/libtheora codec/ogg-vorbis codec/speex communication/im/pidgin etc. You can determine which package group is currently installed on your system: $ pkg list group/system/\* Output would look like: NAME (PUBLISHER) VERSION IFO group/system/solaris-desktop 0.5.11-0.175.0.0.0.0.0 i-- Note that there are not version numbers associated with a group package dependency. The package version that best fits the system will be used, based on other dependencies such as what is listed in incorporation files. Installing a Group To Install a group, simple use the group package name as you would any other package: $ pkg install solaris-small-server  If you want to exclude a package from installing, you can use the --reject flag: $ pkg install --reject audio/audio-utilities solaris-desktop Creating Your Own Group To create your own group package, you can follow the pkg(5) documentation on how to create a package, and use this action for each package that is part of your group:   depend fmri=full/pkg/name type=group

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• Nvidia driver overscan issue second monitor via dvi-d cable

  - by benmichael

  Ok, I know that I have a bit of a bizarre setup, but here goes. I have an old laptop, HP Pavilion 6000. The graphics card in there is a GeForce 7150M. The monitor connection is an old 18pin. The external monitor I use is a Samsung SyncMaster 2333. Don't ask me why, but this monitor only has a dvi-d connection (yes, i have searched it). So I have the monitor plugged into the laptop. If I use any of the Nvidia propriety drivers and try to set the resolution up to 1920x1080 (the monitor's native resolution), I get a massive overscan issue. Over the years I have tried to get this to work, tinkering with my xorg.conf to death. I have also tried this on every Ubuntu since 10.04, on all the corresponding LUbuntus, and on all the Linux Mints since Lisa. Exact same issue. I have even tried it in WinDoze and it works perfectly there (although I did get the error once, but was unable to reproduce it). Using the Open Source drivers it works perfectly iff I switch off the laptop monitor (this makes no difference with the Nvidia drivers). I would have happily gone on using the Open Source drivers, except that since upgrading to LUbuntu 12.10, the Open Source drivers make my monitor completely hazy and have the same overscan issue until I (through the haze, only because I know where things are) go to the monitor settings, activate the laptop's monitor, then deactivate it, and suddenly it comes right. I have to do this every time. So I have to find a way to fix one of them, so I may as well tackle the propriety drivers, hence this overlong question. Amidst other things, I have tried the nvidia-settings, but because it is connected to an 18pin, it detects the monitor as a vga monitor and does not give me overscan correction options. I have tried custom modlines (although there are always more of those try), I have tried using xrandr, and I have tried all the FlatPanelOptions. What I have not tried is a Gentoo build, as I don't have time any more to do that installation, but up to about three years ago when I ran Gentoo exclusively I did not have this issue. Below is an link to an image with a red highlight around the portion of the screen visible to me, the numbers around it are the number of pixels which are cut off. This does seem to drift a few pixels every now and then. Thanks in advance. Nvidia driver issue image

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• Pixels - A cry for some insight

  - by CarrotFile

  I'm pretty new to web developing and I'd love some clarification. Although reading more than one book on the topic, I cannot seem to wrap my head around the pixel concept. I encounter problems with this issue when trying to use CSS and pixel units for design that fits different screen sizes. To my understanding a pixel is the most basic unit used by a monitor in order to compose an image on the screen. So if me resolution is 800 by 600, everything on my screen is rendered using those 800*600 basic building blocks. If I were to enlarge my screen resolution, 3 things would accrue: A. The basic image building block(the pixel) would shrink in size B. The pixels would move close together C. Well, more pixels would now be available All these combined lead to a sharper(depending on the viewing distance) and more detail enabling image. Well so far so good. Here is were I start getting lost: To my knowledge a pixel is not a physical, real object. Monitors are not embedded with a few thousand pixels. I am drawn to this conclusion because anyone can change his screen's resolution, making a pixel on his screen bigger or smaller, and adding or subtracting the amount of total pixels on screen. Adding to that, I have herd that different monitors have different pixel densities. For example Apple's retina monitors. Taking all of the above as my knowledge base, These are my questions: If a pixel has no real world constant size, what does comparing different pixel densities matter? Each screen company can define it's own pixel concept and declare the higher density. What does a bigger pixel density mean? Say we take two screens with the same physical dimensions, but with a different pixel density, am I to assert that the main difference would be the larger density screen being able to display a higher max resolution? Or am I to assert that given the same resolution on both monitors, the higher density one would display a sharper, smaller image? If a pixel is not a fixed size within one monitor, is it a fixed size between the same resolution on two different monitors? For example, would two different monitors, set to the same resolution, be comprised of same size, same quantity pixels? I'd love some help (:

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• HTML5 game programming style

  - by fnx

  I am currently trying learn javascript in form of HTML5 games. Stuff that I've done so far isn't too fancy since I'm still a beginner. My biggest concern so far has been that I don't really know what is the best way to code since I don't know the pros and cons of different methods, nor I've found any good explanations about them. So far I've been using the worst (and propably easiest) method of all (I think) since I'm just starting out, for example like this: var canvas = document.getElementById("canvas"); var ctx = canvas.getContext("2d"); var width = 640; var height = 480; var player = new Player("pic.png", 100, 100, ...); also some other global vars... function Player(imgSrc, x, y, ...) { this.sprite = new Image(); this.sprite.src = imgSrc; this.x = x; this.y = y; ... } Player.prototype.update = function() { // blah blah... } Player.prototype.draw = function() { // yada yada... } function GameLoop() { player.update(); player.draw(); setTimeout(GameLoop, 1000/60); } However, I've seen a few examples on the internet that look interesting, but I don't know how to properly code in these styles, nor do I know if there are names for them. These might not be the best examples but hopefully you'll get the point: 1: Game = { variables: { width: 640, height: 480, stuff: value }, init: function(args) { // some stuff here }, update: function(args) { // some stuff here }, draw: function(args) { // some stuff here }, }; // from http://codeincomplete.com/posts/2011/5/14/javascript_pong/ 2: function Game() { this.Initialize = function () { } this.LoadContent = function () { this.GameLoop = setInterval(this.RunGameLoop, this.DrawInterval); } this.RunGameLoop = function (game) { this.Update(); this.Draw(); } this.Update = function () { // update } this.Draw = function () { // draw game frame } } // from http://www.felinesoft.com/blog/index.php/2010/09/accelerated-game-programming-with-html5-and-canvas/ 3: var engine = {}; engine.canvas = document.getElementById('canvas'); engine.ctx = engine.canvas.getContext('2d'); engine.map = {}; engine.map.draw = function() { // draw map } engine.player = {}; engine.player.draw = function() { // draw player } // from http://that-guy.net/articles/ So I guess my questions are: Which is most CPU efficient, is there any difference between these styles at runtime? Which one allows for easy expandability? Which one is the most safe, or at least harder to hack? Are there any good websites where stuff like this is explained? or... Does it all come to just personal preferance? :)

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• Enterprise with eyes on NoSQL

  - by thegreeneman

  Since joining Oracle a few months back, I have had the fortune of being able to interact with a number of large enterprise organizations and discuss their current state of adoption for NoSQL database technology.   It is worth noting that a large percentage of these organizations do have some NoSQL use and have been steadily increasing their understanding of its applicability for certain data management workloads.   Thru those discussions I’ve learned that it seems one of the biggest issues confronting enterprise adoption of NoSQL databases is the lack of standards for access, administration and monitoring.    This was not so much of an issue with the early adopters of NoSQL technology because they employed a highly DevOps centric approach to application deployment leaving a select few highly qualified developers with the task of managing the production of the system that they designed and implemented. However, as NoSQL technology moves out of the startup and into the hands of larger corporate entities, developers with a broad skill set that are capable of both development and I.T. type production management are in short supply and quickly get moved on to do new projects, often moving to different roles within the company.  This difference in the way smaller more agile startups operate as compared to more established companies is revealing a gap in the NoSQL technology segment that needs to get addressed.    This is one of places that a company such as Oracle has a leg up in the NoSQL Database front.  A combination of having gone thru a past database maturization process,  combined with a vast set of corporate relationships that have grown hand in hand to solve these types of issues, Oracle is in a great place to lead the way in closing the requirements gap for NoSQL technology.  Oracle's understanding of the needs specific to mature organizations have already made their way into the Oracle’s NoSQL Database offering with features such as:  One click cluster deployment with visual topology planning,  standards based monitoring protocols such as SNMP, support for data access for reporting via standard SQL  and integration with emerging standards for data access such as MapReduce.  Given the exciting developments we’re driving in the Oracle NoSQL Database group, I will have a lot more to say about this topic as we move into the second half of the year.

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• Applying DDD principles in a RESTish web service

  - by Andy

  I am developing an RESTish web service. I think I got the idea of the difference between aggregation and composition. Aggregation does not enforce lifecycle/scope on the objects it references. Composition does enforce lifecycle/scope on the objects it contain/own. If I delete a composite object then all the objects it contain/own are deleted as well, while the deleting an aggregate root does not delete referenced objects. 1) If it is true that deleting aggregate roots does not necessary delete referenced objects, what sense does it make to not have a repository for the references objects? Or are aggregate roots as a term referring to what is known as composite object? 2) When you create an web service you will have multiple endpoints, in my case I have one entity Book and another named Comment. It does not make sense to leave the comments in my application if the book is deleted. Therefore, book is a composite object. I guess I should not have a repository for comments since that would break the enforcement of lifecycle and rules that the book class may have. However I have URL such as (examples only): GET /books/1/comments POST /books/1/comments Now, if I do not have a repository for comments, does that mean I have to load the book object and then return the referenced comments? Am I allowed to return a list of Comment entities from the BookRepository, does that make sense? The repository for Book may eventually become rather big with all sorts of methods. Am I allowed to write JPQL (JPA queries) that targets comments and not books inside the repository? What about pagination and filtering of comments. When adding a new comment triggered by the POST endpoint, do you need to load the book, add the comment to the book, and then update the whole book object? What I am currently doing is having a own CommentRepository, even though the comments are deleted with the book. I could need some direction on how to do it correct. Since you are exposing not only root objects in RESTish services I wonder how to handle this at the backend. I am using Hibernate and Spring.

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• Nails vs Screws (C# List vs Dictionary)

  - by MarkPearl

  General This may sound like a typical noob statement, but I’m finding out in a very real way that just because you have a solution to a problem, doesn’t necessarily mean it is the best solution. This was reiterated to me when a friend of mine suggested I look at using Dictionaries instead of Lists for a particular problem – he was right, I have always just assumed that because lists solved my problem I did not need to look elsewhere. So my new manifesto to counter this ageless problem is as follows… Look for a solution that will logically work Once you have a solution look for possible alternatives Decide why your current solution is the best approach compared to the alternatives If it is.. use it till something better comes along, if it isnt…. change What’s the difference between Lists & Dictionaries Both lists and dictionaries are used to store collections of data. Assume we had the following declarations… var dic = new Dictionary<string, long>(); var lst = new List<long>(); long data;   With a list, you simply add the item to the list and it will add the item to the end of the list. lst.Add(data); With a dictionary, you need to specify some sort of key and the data you want to add so that it can be uniquely identified. dic.Add(uniquekey, data);   Because with a dictionary you now have unique identifier, in the background they provide all sort’s of optimized algorithms to find your associated data. What this means is that if you are wanting to access your data it is a lot faster than a List. So when is it appropriate to use either class? For me, if I can guarantee that each item in my collection will have a unique identifier, then I will use Dictionaries instead of Lists as there is a considerable performance benefit when accessing each data item. If I cannot make this sort of guarantee, then by default I will use a list. I know this is all really basic, and I hope I haven’t missed some fundamental principle… If anyone would like to add their 2 cents, please feel free to do so…

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• PASS: FY10 Actuals Posted

  - by Bill Graziano

  Earlier this year we published preliminary fiscal year 2010 financials to the Governance page on the PASS web site.  Please remember that FY10 runs from July 1st, 2009 through June 30th, 2010 and includes the November 2009 Summit.  We do our fiscal year this way so that the Summit falls earlier in the fiscal year.  The financials we had posted were P&L numbers at the portfolio level.  Prior to this we had posted our detailed budget but only posted the auditors report at the end of each year.  Today we updated our published financials to include: Pre-audit actuals from FY10 at the same level as our budget.  The document has both actuals and budget for FY10 side by side.  This is over 20 pages of detailed financial information covering hundreds of line-items. A letter describing key differences between our budget and actuals.  I walked through each line item where the difference was greater than $25,000 and explained what happened and why. We updated the financial graph going back to 2003 to include FY10. This update should “close the loop” on our financials.  You can now start with the published budget and compare it to the finished financials at the same level of detail.  We also plan to publish the auditor’s report when that is completed -- as we do every year. Overall I’m very happy with how FY10 turned out.  Keep in mind that this was the November 2009 Summit so we were still facing economic challenges.  With all that we were roughly break-even showing a $15,000 profit on $3.9 million of revenue.  I didn’t find anything shocking in reviewing our actual vs. budget but there were a few things that needed explanation.  You can see those in the letter on the governance page. Please keep in mind that these are the actuals from our operating financials.  The auditor may have us make adjustments for depreciation or other financial transactions.  We may also account for certain transactions differently for tax purposes than we do for financial reporting purposes.  I feel these financial statements give you the clearest picture of how our organization spends its money. We were late publishing these this year.  We were working through some tax issues and that delayed our ability to file our final tax forms which delayed this process.  In hindsight I should have published these documents as soon as we had them and not waited for the tax issues.  We’ll do this better in the future. And on a final note, you don’t need to login to view these documents.  If you have any questions you can post them here.  If we get more than a few questions we may see about creating some forums for financial issues on the PASS web site.

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• POST and PUT requests – is it just the convention?

  - by bckpwrld

  I've read quite a few articles on the difference between POST and PUT and in when the two should be used. But there are still few things confusing me ( hopefully questions will make some sense ): 1) We should use PUT to create resources when we want clients to specify the URI of the newly created resources and we should use POST to create resources when we let service generate the URI of the newly created resources. a) Is it just by convention that POST create request doesn't contain an URI of the newly created resource or POST create request actually can't contain the URI of the newly created resource? b) PUT has idempotent semantics and thus can be safely used for absolute updates ( ie we send entire state of the resource to the server ), but not also for relative updates ( ie we send just changes to the resource state ), since that would violate its semantics. But I assume it's still possible for PUT to send relative updates to the server, it's just that in that case the PUT update won't be idempotent? 2) I've read somewhere that we should "use POST to append a resource to a collection identified by a service-generated URI". a) What exactly does that mean? That if URIs for the resources were generated by a server ( thus the resources were created via POST ), then ALL subsequent resources should also be created via POST? Thus, in such situation no resource should be created via PUT? b) If my assumption under a) is correct, could you elaborate why we shouldn't create some resources via POST and some via PUT ( assuming server already contains a collection of resources created via POST )? REPLY: 1) Please correct me if I'm wrong, but from your post and from the link you've posted, it seems: a) The Request-URI in POST is interpreted by server as the URI of the service. Thus, it could just as easily be interpreted as an URI of a newly created resource, if server code was written to recognize Request-URI as such b) Similarly, PUT is able to send relative updates, it's just that service code is usually written such that it will complain if PUT updates are relative. 2) Usually, create has fallen into the POST camp, because of the idea of "appending to a collection." It's become the way to append a resource to a list of resources. I don't quite understand the reasoning behind the idea of "appending to a collection" and why this idea prefers POST for create. Namely, if we create 10 resources via PUT, then server will contain a collection of 10 resources and if we then create another resource, then server will append this resource to that collection ( which will now contain 11 resources )?! Uh, this is kinda confusing thank you

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• How do you conquer the challenge of designing for large screen real-estate?

  - by Berin Loritsch

  This question is a bit more subjective, but I'm hoping to get some new perspective. I'm so used to designing for a certain screen size (typically 1024x768) that I find that size to not be a problem. Expanding the size to 1280x1024 doesn't buy you enough screen real estate to make an appreciable difference, but will give me a little more breathing room. Basically, I just expand my "grid size" and the same basic design for the slightly smaller screen still works. However, in the last couple of projects my clients were all using 1080p (1920x1080) screens and they wanted solutions to use as much of that real estate as possible. 1920 pixels across provides just under twice the width I am used to, and the wide screen makes some of my old go to design approaches not to work as well. The problem I'm running into is that when presented with so much space, I'm confronted with some major problems. How many columns should I use? The wide format lends itself to a 3 column split with a 2:1:1 split (i.e. the content column bigger than the other two). However, if I go with three columns what do I do with that extra column? How do I make efficient use of the screen real estate? There's a temptation to put everything on the screen at once, but too much information actually makes the application harder to use. White space is important to help make sense of complex information, but too much makes related concepts look too separate. I'm usually working with web applications that have complex data, and visualization and presentation is key to making sense of the raw data. When your user also has a large screen (at least 24"), some information is out of eye sight and you need to move the pointer a long distance. How do you make sure everything that's needed stays within the visual hot points? Simple sites like blogs actually do better when the width is constrained, which results in a lot of wasted real estate. I kind of wonder if having the text box and the text preview side by side would be a big benefit for the admin side of that type of screen? (1:1 two column split). For your answers, I know almost everything in design is "it depends". What I'm looking for is: General principles you use How your approach to design has changed I'm finding that i have to retrain myself how to work with this different format. Every bump in resolution I've worked through to date has been about 25%: 640 to 800 (25% increase), 800 to 1024 (28% increase), and 1024 to 1280 (25% increase). However, the jump from 1280 to 1920 is a good 50% increase in space--the equivalent from jumping from 640 straight to 1024. There was no commonly used middle size to help learn lessons more gradually.

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• How is a switch statement better than a series of if statements? [closed]

  - by user1276078

  Possible Duplicate: Should I use switch statements or long if…else chains? I'm working on a small program that will conduct an Insertion Sort. A number will be inputted through the keyboard and stored in a variable I called "num." I've decided to use a switch statement in order to obtain the number inputted. switch( e.getKeyCode() ) { case KeyEvent.VK_0: num = 0; break; case KeyEvent.VK_1: num = 1; break; case KeyEvent.VK_2: num = 2; break; case KeyEvent.VK_3: num = 3; break; case KeyEvent.VK_4: num = 4; break; case KeyEvent.VK_5: num = 5; break; case KeyEvent.VK_6: num = 6; break; case KeyEvent.VK_7: num = 7; break; case KeyEvent.VK_8: num = 8; break; case KeyEvent.VK_9: num = 9; break; } I realized one other course of action could have been to use a set of if statements. if( e.getKeyCode() == KeyEvent.VK_0 ) num = 0; else if( e.getKeyCode() == KeyEvent.VK_1 ) num = 1; etc. for every number up until 9. I then wondered what the essential difference is between a switch statement and a series of if statements. I know it saves space and time to write, but it's not that much. So, my question is, aside from the space, does a switch statement differ from a series of if statments in any way? Is it faster, less error-prone, etc.? This question really doesn't affect my code that much. I was just wondering. Also, this question pertains to the JAVA language, not any other programming language.

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• Correct permissions for /var/www and wordpress

  - by dpbklyn

  Hello and thank you in advance! I am relatively new to ubuntu, so please excuse the newbie-ness of this question... I have set up a LAMP server (ubuntu server 11.10) and I have access via SSH and to the "it works" page from a web browser from inside my network (via ip address) and from outside using dyndns. I have a couple of projects in development with some outside developers and I want to use this server as a development server for testing and for client approvals. We have some Wordpress projects that sit in subdirectories in /var/www/wordpress1 /var/www/wordpress2, etc. I cannot access these sub directories from a browser in order to set up WP--or (I assume) to see the content on a browser. I get a 403 Forbidden error on my browser. I assume that this is a permissions problem. Can you please tell me the proper settings for the permissions to: 1) Allow the developers and me to read/write. 2) to allow WP set up and do its thing 3) Allow visitors to access the site(s) via the web. I should also mention that the subfolder are actually simlinks to folder on another internal hdd--I don't think this will make a difference, but I thought I should disclose. Since I am a newbie to ubuntu, step-by-step directions are greatly appreciated! Thank you for taking the time! dp total 12 drwxr-xr-x 2 root root 4096 2012-07-12 10:55 . drwxr-xr-x 13 root root 4096 2012-07-11 20:02 .. lrwxrwxrwx 1 root root 43 2012-07-11 20:45 admin_media -> /root/django_src/django/contrib/admin/media -rw-r--r-- 1 root root 177 2012-07-11 17:50 index.html lrwxrwxrwx 1 root root 14 2012-07-11 20:42 media -> /hdd/web/media lrwxrwxrwx 1 root root 18 2012-07-12 10:55 wordpress -> /hdd/web/wordpress Here is the result of using chown -R www-data:www-data /var/www total 12 drwxr-xr-x 2 www-data www-data 4096 2012-07-12 10:55 . drwxr-xr-x 13 root root 4096 2012-07-11 20:02 .. lrwxrwxrwx 1 www-data www-data 43 2012-07-11 20:45 admin_media -> /root/django_src/django/contrib/admin/media -rw-r--r-- 1 www-data www-data 177 2012-07-11 17:50 index.html lrwxrwxrwx 1 www-data www-data 14 2012-07-11 20:42 media -> /hdd/web/media lrwxrwxrwx 1 www-data www-data 18 2012-07-12 10:55 wordpress -> /hdd/web/wordpress I am still unable to access via browser...

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• C Programming arrays, I dont understand how I would go about making this program, If anyone can just guide me through the basic outline please :) [on hold]

  - by Rashmi Kohli

  Problem The temperature of a car engine has been measured, from real-world experiments, as shown in the table and graph below: Time (min) Temperature (oC) 0 20 1 36 2 61 3 68 4 77 5 110 Use linear regression to find the engine’s temperature at 1.5 minutes, 4.3 minutes, and any other time specified by the user. Background In engineering, many times we measure several data points in an experiment, but then we need to predict a value that we have not measured which lies between two measured values, such as the problem statement above. If the relation between the measured parameters seems to be roughly linear, then we can use linear regression to find the relationship between those parameters. In the graph of the problem statement above, the relation seems to be roughly linear. Hence, we can apply linear regression to the above problem. Assuming y {y0, y1, …yn-1} has a linear relation with x {x0, x1, … xn-1}, we can say that: y = mx+b where m and b can be found with linear regression as follows: For the problem in this lab, using linear regression gives us the following line (in blue) compared to the measured curve (in red). As you can see, there is usually a difference between the measured values and the estimated (predicted) values. What linear regression does is to minimize those differences and still give us a straight line (blue). Other methods, such as non-linear regression, are also possible to achieve higher accuracy and better curve fitting. Requirements Your program should first print the table of the temperatures similar to the way it’s printed in the problem statement. It should then calculate the temperature at minute 1.5 and 4.3 and show the answers to the user. Next, it should prompt the user to enter a time in minutes (or -1 to quit), and after reading the user’s specified time it should give the value of the engine’s temperature at that time. It should then go back to the prompt. Hints •Use a one dimensional array to store the temperature values given in the problem statement. •Use functions to separate tasks such as calculating m, calculating b, calculating the temperature at a given time, printing the prompt, etc. You can then give your algorithm as well as you pseudo code per function, as opposed to one large algorithm diagram or one large sequence of pseudo code.

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• Advantages to Multiple Methods over Switch

  - by tandu

  I received a code review from a senior developer today asking "By the way, what is your objection to dispatching functions by way of a switch statement?" I have read in many places about how pumping an argument through switch to call methods is bad OOP, not as extensible, etc. However, I can't really come up with a definitive answer for him. I would like to settle this for myself once and for all. Here are our competing code suggestions (php used as an example, but can apply more universally): class Switch { public function go($arg) { switch ($arg) { case "one": echo "one\n"; break; case "two": echo "two\n"; break; case "three": echo "three\n"; break; default: throw new Exception("Unknown call: $arg"); break; } } } class Oop { public function go_one() { echo "one\n"; } public function go_two() { echo "two\n"; } public function go_three() { echo "three\n"; } public function __call($_, $__) { throw new Exception("Unknown call $_ with arguments: " . print_r($__, true)); } } Part of his argument was "It (switch method) has a much cleaner way of handling default cases than what you have in the generic __call() magic method." I disagree about the cleanliness and in fact prefer call, but I would like to hear what others have to say. Arguments I can come up with in support of Oop scheme: A bit cleaner in terms of the code you have to write (less, easier to read, less keywords to consider) Not all actions delegated to a single method. Not much difference in execution here, but at least the text is more compartmentalized. In the same vein, another method can be added anywhere in the class instead of a specific spot. Methods are namespaced, which is nice. Does not apply here, but consider a case where Switch::go() operated on a member rather than a parameter. You would have to change the member first, then call the method. For Oop you can call the methods independently at any time. Arguments I can come up with in support of Switch scheme: For the sake of argument, cleaner method of dealing with a default (unknown) request Seems less magical, which might make unfamiliar developers feel more comfortable Anyone have anything to add for either side? I'd like to have a good answer for him.

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• WiFi on Ubuntu 12.04 custom: downloading unbearably slow

  - by Mark

  iwconfig reports 11 Mbps, yet I've seen as low as <1 KBps. This is the latest in my laundry list of Ubuntu problems in a dual-boot machine (cyberpowerpc custom, intel i7-3820, nvidia gtx 570). I received it two days ago, Windows 7 running fine, still having problems with Ubuntu. The browsing is intermittent but unacceptable. e.g. I could get to this site last night but I couldn't post this question. The downloading is unbearably slow, I can't download anything or install any packages because the speed is so slow. e.g. I am trying to install vim which is inexplicably missing from my 12.04 install (add another one to the problems list) and my download speed reported in the terminal was 241 B/s. Yes, bytes. iwconfig reports 11 Mbps, which further adds to the confusion. User@ubuntu:~$ iwconfig lo no wireless extensions. wlan0 IEEE 802.11bgn ESSID:"linksys" Mode:Managed Frequency:2.437 GHz Access Point: 00:18:39:76:2C:A1 Bit Rate=11 Mb/s Tx-Power=20 dBm Retry long limit:7 RTS thr:off Fragment thr:off Power Management:off Link Quality=36/70 Signal level=-74 dBm Rx invalid nwid:0 Rx invalid crypt:0 Rx invalid frag:0 Tx excessive retries:54 Invalid misc:18 Missed beacon:0 eth0 no wireless extensions. Any ideas? I see this is a problem a lot of people, but none of the on line solutions have worked for me so far. e.g. one site recommends editing the ath9k.conf file in /etc/modprobe.d, yet this file isn't even in the folder: User@ubuntu:/$ cd etc/modprobe.d User@ubuntu:/etc/modprobe.d$ ls alsa-base.conf blacklist-oss.conf blacklist-ath_pci.conf blacklist-rare-network.conf blacklist.conf blacklist-watchdog.conf blacklist-firewire.conf dkms.conf blacklist-framebuffer.conf nvidia-current_hybrid.conf blacklist-modem.conf nvidia-graphics-drivers.conf I think the nvidia gpu might be mucking things up. I had the "blinking cursor" problem when installing in the first place, and then I had the monitor out of range problem as well. I have my faithful Asus laptop, which is running Ubuntu 12.04 just fine. The only difference is executing host -t SOA local in the terminal gives User@ubuntu:~$ host -t SOA local local has SOA record local. nobody.localhost. 42 86400 43200 604800 10800 in my new machine, and the command reports Host local. not found in the laptop. Help would be most welcome, as I am in danger of reverting back to Windows. I'm seriously considering it. Sorry for the length, trying to show my effort in resolving the issue and include terminal snippets that might be helpful.

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• The term "interface" in C++

  - by Flexo

  Java makes a clear distinction between class and interface. (I believe C# does also, but I have no experience with it). When writing C++ however there is no language enforced distinction between class and interface. Consequently I've always viewed interface as a workaround for the lack of multiple inheritance in Java. Making such a distinction feels arbitrary and meaningless in C++. I've always tended to go with the "write things in the most obvious way" approach, so if in C++ I've got what might be called an interface in Java, e.g.: class Foo { public: virtual void doStuff() = 0; ~Foo() = 0; }; and I then decided that most implementers of Foo wanted to share some common functionality I would probably write: class Foo { public: virtual void doStuff() = 0; ~Foo() {} protected: // If it needs this to do its thing: int internalHelperThing(int); // Or if it doesn't need the this pointer: static int someOtherHelper(int); }; Which then makes this not an interface in the Java sense anymore. Instead C++ has two important concepts, related to the same underlying inheritance problem: virtual inhertiance Classes with no member variables can occupy no extra space when used as a base "Base class subobjects may have zero size" Reference Of those I try to avoid #1 wherever possible - it's rare to encounter a scenario where that genuinely is the "cleanest" design. #2 is however a subtle, but important difference between my understanding of the term "interface" and the C++ language features. As a result of this I currently (almost) never refer to things as "interfaces" in C++ and talk in terms of base classes and their sizes. I would say that in the context of C++ "interface" is a misnomer. It has come to my attention though that not many people make such a distinction. Do I stand to lose anything by allowing (e.g. protected) non-virtual functions to exist within an "interface" in C++? (My feeling is the exactly the opposite - a more natural location for shared code) Is the term "interface" meaningful in C++ - does it imply only pure virtual or would it be fair to call C++ classes with no member variables an interface still?

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• Deleted Myself from Admin Group - Now Getting Error usermod: cannot lock /etc/passwd; try again later

  - by BubbaJ

  I have a laptop with Ubuntu 11.10 that is shared between myself and two other family members. My user id was setup as the only "Administrator" on the laptop. The other users were setup as "Standard" users. In my attempt to try to add myself to the user groups for the other users, I somehow deleted myself from the admin groups. I used the "usermod" command from the terminal. I must have neglected to include the proper switches or syntax for the update. It looks like I successfully added my userid to the group associated with my wife's account. When I use the "groups" command, I can see only my id and my wife's id in the list. I no longer see the "admin" or "adm" groups, and others that used to be listed. When I go into System Settings User Accounts it looks like my ID is now listed as a "Standard" user. I would like to change my account back to "Administrator", but now I can't. I did some searches for solutions and found that I would need to boot into Recovery Mode and execute the usermod command from the root session. I was able to successfully boot into Recovery Mode and get to the root session. I was trying to execute the command "usermod -a -G admin user1" to add my id (user1) back to the admin group. When I execute the command from the root session, I get the error message "usermod: cannot lock /etc/passwd; try again later". I tried preceding the usermod command with "sudo", but it didn't make a difference, same error. I then tried adding a new user using adduser, thinking I would try to create a new userid and make the new userid part of the admin group. I get the same error using the adduser command. I saw some posts that recommend looking for and deleting files that end in ".lock" in the etc directory. The only file I found was .pwd.lock which I haven't touched. I am at a loss as to what to try next. I am relatively inexperienced with Ubuntu and Linux, so alot of this is new to me. Any help you can provide would be much appreciated.

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• How do I render from one render target to another?

  - by Chaotikmind

  I have two render targets: a fake backbuffer; a special render target where I do all my rendering. a light render target; where I render my light fx. I'm sure I'm rendering correctly on both. The problem arises when I overlay the light render target onto the fake backbuffer by drawing a quad covering it: DxEngine.DrawSprite(0.0f, 0.0f, 0.0f, (float)DxEngine.GetWidth(), (float)DxEngine.GetHeight(), 0xFFFFFFFF, LightSurface->GetTexture()); Regardless of what's in the light target, nothing is rendered onto the other target. I tried clearing the light target with full-white or full-black, but still get nothing. Fake backbuffer created with Direct3dDev->CreateTexture(Width, Height, 1, D3DUSAGE_RENDERTARGET, D3DFMT_X8R8G8B8, D3DPOOL_DEFAULT, &Texture, nullptr); Light render target created with Direct3dDev->CreateTexture(Width, Height, 1, D3DUSAGE_RENDERTARGET, D3DFMT_A8R8G8B8, D3DPOOL_DEFAULT, &Texture, nullptr); I also tried to create both with D3DFMT_A8R8G8B8, again without difference. Both targets have the same width and height. Only the fixed pipeline is used DirectX setup for rendering : Direct3dDev->SetSamplerState(0, D3DSAMP_MINFILTER, D3DTEXF_LINEAR ); Direct3dDev->SetSamplerState(0, D3DSAMP_MAGFILTER, D3DTEXF_LINEAR ); Direct3dDev->SetSamplerState(0, D3DSAMP_MIPFILTER, D3DTEXF_NONE ); Direct3dDev->SetSamplerState(0, D3DSAMP_ADDRESSU, D3DTADDRESS_WRAP ); Direct3dDev->SetSamplerState(0, D3DSAMP_ADDRESSV, D3DTADDRESS_WRAP ); Direct3dDev->SetRenderState(D3DRS_CULLMODE, D3DCULL_NONE); Direct3dDev->SetRenderState(D3DRS_LIGHTING, false); Direct3dDev->SetRenderState(D3DRS_ZENABLE, D3DZB_TRUE); Direct3dDev->SetRenderState(D3DRS_ZWRITEENABLE,D3DZB_TRUE); Direct3dDev->SetRenderState(D3DRS_ZFUNC,D3DCMP_LESSEQUAL); Direct3dDev->SetRenderState(D3DRS_ALPHABLENDENABLE, true ); Direct3dDev->SetRenderState(D3DRS_ALPHAREF, 0x00000000ul); Direct3dDev->SetRenderState(D3DRS_ALPHATESTENABLE, true); Direct3dDev->SetRenderState(D3DRS_ALPHAFUNC,D3DCMP_GREATER); Direct3dDev->SetRenderState(D3DRS_SRCBLEND, D3DBLEND_SRCALPHA ); Direct3dDev->SetRenderState(D3DRS_DESTBLEND, D3DBLEND_INVSRCALPHA ); Direct3dDev->SetTextureStageState(0, D3DTSS_COLORARG1, D3DTA_TEXTURE); Direct3dDev->SetTextureStageState(0, D3DTSS_COLORARG2, D3DTA_DIFFUSE); Direct3dDev->SetTextureStageState(0, D3DTSS_COLOROP, D3DTOP_MODULATE); Direct3dDev->SetTextureStageState(0, D3DTSS_ALPHAARG1, D3DTA_TEXTURE); Direct3dDev->SetTextureStageState(0, D3DTSS_ALPHAARG2, D3DTA_DIFFUSE); Direct3dDev->SetTextureStageState(0, D3DTSS_ALPHAOP, D3DTOP_MODULATE); Direct3dDev->SetTextureStageState(0, D3DTSS_RESULTARG, D3DTA_CURRENT); Direct3dDev->SetTextureStageState(0, D3DTSS_TEXCOORDINDEX, D3DTSS_TCI_PASSTHRU); //ensure the first stage is not used for now Direct3dDev->SetTextureStageState(1, D3DTSS_COLOROP, D3DTOP_DISABLE); How can I do this right?

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• Hyper-V Server 2012 - "No Active Network Adapter Found"

  - by Vazgen

  I just installed Hyper-V Server 2012 and first thing I get when logging in to the host is "No Active Network Adapter Found". I know I have to use the following command: pnputil –i –a <inf file> I found the Realtek Network Driver for my MB for Windows 7 here: http://za.asus.com/Motherboards/AMD_AM3Plus/M5A97_PRO/#download But there is not one for Windows 8 so I tried using the Win 7 one with PnPUtil, and I got: Failed to install the driver on any of the devices on the system: No more data is available. Does this really mean I can't use Hyper-V Server 2012 with my motherboard? This is hard to believe because I installed Windows Server 2012 directly (prior to installing Hyper-V Server 2012) on this same computer and I had internet access on that server without even configuring any drivers. Ok so I went to PowerShell from cmd and typed Get-NetIPConfiguration And it said the name of my driver (Realtek) and "Disconnected". I switched the Ethernet wire from my primary computer with my server and everything was fixed. Interestingly, my primary computer connects to the internet just fine with the wire that was previously in the server, just with much slower speed. Does anyone know what could be the difference between these 2 Ethernet wires?

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• Hyper-V Virtual Ethernet Adapter not working probperly - Code 31 on Host

  - by Chris S

  I have a Windows 8.1 machine with Hyper-V installed. From a "clean" Hyper-V configuration I open Virtual Switch Manager, create an External switch with "Allow management operating system to share this network adapter", click OK. Everything seems to work properly but the host loses network connectivity. Opening Device Manager, the "Hyper-V Virtual Ethernet Adapter" is shown under the Network Adapter section with a yellow triangle, and the following message: This device is not working properly because Windows cannot load the drivers required for this device. (Code 31) An object ID was not found in the file. I tried "Uninstalling" the device, seems to work, but the device doesn't actually remove. I tried removing and re-adding the Hyper-V feature completely, no difference. Tried scf /scannow, no problems. System and Application logs show no errors. The Hyper-V-VMMS Networking log shows the following: Log Name: Microsoft-Windows-Hyper-V-VMMS-Networking Source: Microsoft-Windows-Hyper-V-VMMS Date: 10/24/2013 10:53:07 AM Event ID: 26088 Description: Failed to apply static IP settings to internal Ethernet adapter {A813DE9A-BE70-4FAE-AD31-BE4D54505A4B} ('885435B8-BE65-4EE9-826D-AB56035237ED'): Unspecified error (0x80004005). If I try to remove the Virtual Switch in Hyper-V Virtual Switch Manager I get this in that same log: Log Name: Microsoft-Windows-Hyper-V-VMMS-Networking Source: Microsoft-Windows-Hyper-V-VMMS Date: 10/24/2013 11:19:47 AM Event ID: 26142 Description: Failed while removing virtual Ethernet switch. Trying to remove the Virtual Switch leads to an error: Error applying Virtual Switch Properties changes Failed while removing virtual Ethernet switch. VM Networking does work.

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