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  • Infinite detail inside Perlin noise procedural mapping

    - by Dave Jellison
    I am very new to game development but I was able to scour the internet to figure out Perlin noise enough to implement a very simple 2D tile infinite procedural world. Here's the question and it's more conceptual than code-based in answer, I think. I understand the concept of "I plug in (x, y) and get back from Perlin noise p" (I'll call it p). P will always be the same value for the same (x, y) (as long as the Perlin algorithm parameters haven't changed, like altering number of octaves, et cetera). What I want to do is be able to zoom into a square and be able to generate smaller squares inside of the already generated overhead tile of terrain. Let's say I have a jungle tile for overhead terrain but I want to zoom in and maybe see a small river tile that would only be a creek and not large enough to be a full "big tile" of water in the overhead. Of course, I want the same net effect as a Perlin equation inside a Perlin equation if that makes sense? (aka. I want two people playing the game with the same settings to get the same terrain and details every time). I can conceptually wrap my head around the large tile being based on an "zoomed out" coordinate leaving enough room to drill into but this approach doesn't make sense in my head (maybe I'm wrong). I'm guessing with this approach my overhead terrain would lose all of the cohesiveness delivered by the Perlin. Imagine I calculate (0, 0) as overhead tile 1 and then to the east of that I plug in (50, 0). OK, great, I now have 49 pixels of detail I could then "drill down" into. The issue I have in my head with this approach (without attempting it) is that there's no guarantee from my Perlin noise that (0,0) would be a good neighbor to (50,0) as they could have wildly different "elevations" or p/resultant values returning from the Perlin equation when I generate the overhead map. I think I can conceive of using the Perlin noise for the overhead tile to then reuse the p value as a seed for the "detail" level of noise once I zoom in. That would ensure my detail Perlin is always the same configuration for (0,0), (1,0), etc. ad nauseam but I'm not sure if there are better approaches out there or if this is a sound approach at all.

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  • Culture Shmulture?

    - by steve.diamond
    I've been thinking about "Customer Experience Management" lately. Here at Oracle, we arguably have the most complete suite of applications for managing the customer experience across and in the context of multiple channels -- from marketing to loyalty to contact center to self-service to analytics offerings, and more. And stay tuned, because in coming months let's just say we'll have even more to talk about on this front. But that said............ Last weekend my wife and I stayed at one of the premiere hotel chains on the planet. I won't name them, but we all know the short list. It could have been the St. Regis or the Ritz Carlton or Four Seasons or Hyatt Park or....This stay, at this particular hotel, was simply outstanding. Within a chain known for providing "above and beyond" levels of service, this particular hotel, under this particular manager, exceeded expectations on so many fronts. For example, at the Spa we mentioned to the two attendants that my wife is seven months pregnant and that we had previously had a lot of trouble conceiving. We then went to our room. Ten minutes later we heard a knock at the door and received a plate of chocolate covered strawberries with a heartfelt note and an inspiring quote, signed by the two spa attendees. The following day we arranged to have a bellhop drive us to the beach. Although they had a pre-arranged beach shuttle service with time limits, etc., he greeted us by saying, "I'm yours for the day until 4 p.m. Whatever you want to do is fine by me, as long as it's legal!" The morning that we left we arranged to have a taxi drive us to the airport--a nearly 40 mile drive. What showed up was a private coach complete with navy blue suited driver dude. And we were charged the taxi fare price. And there were many other awesome exchanges I won't mention here, although I did email the GM of this hotel two nights ago and expressed our effusive praise and gratitude. I'd submit that this hotel chain would have a definitive advantage using even more Oracle software to manage and optimize its customer interactions (yes, they are a customer). But WITHOUT the culture--that management team--and that instillation of aligned values across all employees of exemplifying 'the golden rule,' I wonder how much technology really matters in providing a distinctively positive and memorable customer experience. Lest you think I'm alone in these pontifications, have you read Paul Greenberg's blog lately? Have you seen one of his most recent posts? Now this SPECIFIC post is NOT about customer service per se. But it is about people. So yes, please think long and hard about the technology you seek to deploy. But never forget who will be interacting with your systems, and your customers.

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  • Understanding the JSF Lifecycle and ADF Optimized Lifecycle

    - by Steven Davelaar
    While coaching ADF development teams over the years, I have noticed that many developers lack a basic understanding of Java Server Faces, in particular the JSF lifecycle and how ADF optimizes this lifecycle in specific situations. As a result, ADF developers who are tasked to build a seemingly simple ADF page, can get extremely frustrated by the -in their eyes- unexpected or unlogical behavior of ADF.  They start to play with the immediate property and the partialTriggers property in a trial-and-error manner. Often, they play with these properties until their specific issue is solved, unaware of other more severe bugs that might be introduced by the values they choose for these properties. So, I decided to submit a presentation for the UKOUG entitled "What you need to know about JSF to be succesful with ADF".  The abstract was accepted, and I started putting together the presentation and demo application. I built up a demo application step-by-step, trying to cover the JSF-related  top issues and challenges I encountered over the years in a simple "Hello World" demo. This turned out to be both a very time-consuming and very interesting journey. I had never thought I would learn so much myself in preparing this presentation. I never thought I would end up with potentially controversial conclusions like "Never set immediate=true on an editable component".  I did not realize the sometimes immense implications of the ADF optimized lifecycle beforehand. I never thought that "Hello World" demo's could get so complex. But as I went on I was confident this was valuable material, even for experienced ADF developers with a good understanding of JSF. When I finished, I realized the original title and abstract was misleading, as was the target audience. Yes, it was covering the JSF lifecycle, but no other aspects of JSF you need to know for ADF development. Yes, it was covering some JSF basics as mentioned in the abstract, but all in all it had become a pretty advanced presentation. At the same time, the issues discussed are very common, novice ADF developers might easily run into them while building their first pages. I ran out of time, so I decided to just present what I had, apologizing at the beginning for the misleading title, showing a second slide with a better title "18 invaluable lessons about ADF-JSF interaction". I think the presentation was well received overall, although people who don't like it or don't understand it, usually don't come and tell you afterwards.... I am still struggling with the title, for this blog post I used yet another title, anyway, you can download the presentation-that-still-lacks-a-good-title here. The finished JDev 11.1.1.6 demo app can be downloaded here.  The 18 lessons mentioned in the presentation are summarized here. As mentioned on the last slide, print out the lessons, and learn them by heart, I am pretty sure it will save you lots of time and frustration!

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  • Open the SQL Server Error Log with PowerShell

    - by BuckWoody
    Using the Server Management Objects (SMO) library, you don’t even need to have the SQL Server 2008 PowerShell Provider to read the SQL Server Error Logs – in fact, you can use regular old everyday PowerShell. Keep in mind you will need the SMO libraries – which can be installed separately or by installing the Client Tools from the SQL Server install media. You could search for errors, store a result as a variable, or act on the returned values in some other way. Replace the Machine Name with your server and Instance Name with your instance, but leave the quotes, to make this work on your system: [reflection.assembly]::LoadWithPartialName("Microsoft.SqlServer.Smo") $machineName = "UNIVAC" $instanceName = "Production" $sqlServer = new-object ("Microsoft.SqlServer.Management.Smo.Server") "$machineName\$instanceName" $sqlServer.ReadErrorLog() Want to search for something specific, like the word “Error”? Replace the last line with this: $sqlServer.ReadErrorLog() | where {$_.Text -like "Error*"} Script Disclaimer, for people who need to be told this sort of thing: Never trust any script, including those that you find here, until you understand exactly what it does and how it will act on your systems. Always check the script on a test system or Virtual Machine, not a production system. Yes, there are always multiple ways to do things, and this script may not work in every situation, for everything. It’s just a script, people. All scripts on this site are performed by a professional stunt driver on a closed course. Your mileage may vary. Void where prohibited. Offer good for a limited time only. Keep out of reach of small children. Do not operate heavy machinery while using this script. If you experience blurry vision, indigestion or diarrhea during the operation of this script, see a physician immediately. Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • Copying & Pasting Rows Between Grids in SQL Developer

    - by thatjeffsmith
    Apologies for slacking on the blogging front here lately. Still mentally hung over from Open World, and lots of things going on behind the scenes here in Oracle-land. Whilst (love that word) blogging is part of my job, it’s not the ONLY part of my job So a super-quick and dirty ‘trick’ this morning. Copying Query Result Record as New Row in Table Copy and paste is something everyone ‘gets.’ I don’t know we have to thank for that, whether it’s Microsoft or Xerox, but it’s been ingrained in our way of dealing with all things computers. Almost to the detriment of some of our users – they’ll use Copy and Paste when perhaps our Export feature is superior, but I digress. Where it does work just fine is when you want to create a new row in your table that matches a row you have retrieved from an executed query. Just click in the gutter or row number to get the entire row selected Once you have your data selected, do your thing, i.e. ctrl+C or Command/Apple+C or whatever. Now open your view or table editor, go to the data page, and ask for a new row. New record, no data Paste in the data from the clipboard. It’s smart enough to paste the separate values out to the separate columns. The clipboard saves the day, again. If your columns orders are different, just change the order in the grids. If you have extra information, don’t copy the entire row. I know, I know – Jeff this is too simple, why are you wasting our time here? It seems intuitive, but how many of you actually tried this before reading it just now? I seem to get more positive feedback from the very basic user interface 101 tips than the esoteric click-click-click-ctrl-shift-click tricks I prefer to post. Lots of interesting stuff on tap, so stay tuned!

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  • Running Apache and Tomcat together on different subdomains?

    - by Ritesh M Nayak
    Posted this on ServerFault but didn't get a response. Hoping I will have better luck on the Ubuntu site. I have been trying to get this working the whole of today. I have a server which resolves to the domain example.com . This is running Apache2 and Tomcat 6. The requirement is to direct requests to example.com to apache2 and app.example.com to Tomcat. I know I have to do a VirtualHost proxy pass for this to work. Here are the settings on my server. /etc/hosts file looks something like this 127.0.0.1 localhost localhost.localdomain example.com app.example.com I have two virtual host files for the different domains in /etc/apache2/sites-enabled /etc/apache2/sites-enabled/example.com looks like this <VirtualHost *:80> # Admin email, Server Name (domain name) and any aliases ServerAdmin webmaster@localhost ServerName example.com ServerAlias www.example.com DocumentRoot /var/www <Directory /> Options FollowSymLinks AllowOverride None </Directory> <Directory /var/www/> Options Indexes FollowSymLinks MultiViews AllowOverride None Order allow,deny allow from all </Directory> ScriptAlias /cgi-bin/ /usr/lib/cgi-bin/ <Directory "/usr/lib/cgi-bin"> AllowOverride None Options +ExecCGI -MultiViews +SymLinksIfOwnerMatch Order allow,deny Allow from all </Directory> ErrorLog /var/log/apache2/error.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn CustomLog /var/log/apache2/access.log combined Alias /doc/ "/usr/share/doc/" <Directory "/usr/share/doc/"> Options Indexes MultiViews FollowSymLinks AllowOverride None Order deny,allow Deny from all Allow from 127.0.0.0/255.0.0.0 ::1/128 </Directory> </VirtualHost> /etc/apache2/sites-enabled/app.example.com file looks like this <VirtualHost *:80> ServerName app.example.com ServerAlias www.app.example.com ProxyPreserveHost On ProxyPass / http://localhost:8080/ ProxyPassReverse / http://localhost:8080/ SetEnv force-proxy-request-1.0 1 SetEnv proxy-nokeepalive 1 </VirtualHost> mod_proxy and mod_rewrite are both enabled on the apache instance. I have a CNAME entry for both example.com and app.example.com. When accessing app.example.com, I get an 403 forbidden, saying I have no access to / on the server. What am I doing wrong?

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  • Cannot dual Windows XP and Ubuntu

    - by Fabio Machado
    I am new to Ubuntu and at the moment I am trying to get Ubuntu 12.10 to one of my machines. The machine is a Pentium 4 @ 3.06, 2Gb RAM, 200GB Hard Drive and a NVidia GeForce 8800 GT. A few days ago, I tried Ubuntu without installing and it worked perfectly. Yesterday, I decided to formatted the hard drive and divide my hard drive into four partitions: 1 for the XP, 1 for Ubuntu, 1 for swamp and 1 where I will have my documents. Everything went great, I installed XP and then Ubuntu but I did something wrong on the partition window (Ubunto partion window) that I ended up without boot loader. This morning, I formatted everything again, installed XP and when I went to install Ubuntu (with the same DVD as before) the problems started. First, I had a black screen with a msg written with white text saying something like: unable to find a medium containing a live file system. After I burned another CD and tried again, I got stuck at the red dots (loading screen). I then went online and I read somewhere that it could be the CD, so I checked the integrity of the CD and everything was fine. I also unplugged all USBs connected to the computer and nothing changed. I goggled further options to try to solve my problem and some users suggested that people having these types of problems should try the alternate installation, which if I am not wrong is for networks. I then tried to install and yes the installation process was different from the normal CD, but it did get stuck on a page where it was doing something, like: ...finding ethd0 and it was stuck on the 100%. I tried USB installation as well and it also got stuck at the red dots (I do not have USB 3.0 on the computer in question). I have burned 5 different CD's and all at low speed. I checked the integrity and all are fine. I downloaded other distribution as well as other versions of Ubuntu and I still cannot install or even run the Live CD of Ubuntu or any other distribution. What is really annoying me is that everything was working perfectly before, when I first tried to install Ubuntu. Anyway, any help is welcome. Edit: My boot load is normal, no errors and all the hardware is working fine. I forgot to mention that after the loading screen (red dots) gets stuck, the DVD drive and the hard drive goes into idle state. I also restored the default values of the BIOS and no luck.

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  • Binary on the Coat of Arms of the Governor General of Canada

    - by user132636
    Can you help me further this investigation? Here is about 10% of the work I have done on it. I present it only to see if there are any truly curious people among you. I made a video a few weeks ago showing some strange things about the Governor General's Coat of Arms and the binary on it. Today, I noticed something kinda cool and thought I would share. Here is the binary as it appears on the COA: 110010111001001010100100111010011 As DEC: 6830770643 (this is easily found on the web) Take a close look at that number. What do you notice about it? It has a few interesting features, but here is the one no one has pointed out... Split it down the middle and you have 68307 70643. The first digit is double the value of the last digit. The second digit is double the second last digit. The third digit is half of the third to last digit. And the middle ones are even or neutral. At first, I thought of it as energy. ++-nnnn+-- But actually you can create something else with it using the values. 221000211. See how that works. You may be asking why that is significant. Bare with me. I know 99% are rolling their eyes. 221000211 as base3 gives you this as binary: 100011101000111 100011101000111 as HEX is 4747, which converts to "GG". Initials of Governor General. GG.ca is his website. When you convert to base 33 (there are 33 digits in the original code) you get "GOV" Interesting? :D There is a lot more to it. I'll continue to show some strange coincidences if anyone is interested. Sorry if I am not explaining this correctly. By now you have probably figured out that I have no background in this. Which is why I am here. Thank you.

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  • Question on design of current pagination implementations

    - by Freshblood
    I have checked pagination implementations on asp.net mvc specifically and i really feel that there is something less efficient in implementations. First of all all implementations use pagination values like below. public ActionResult MostPopulars(int pageIndex,int pageSize) { } The thing that i feel wrong is pageIndex and pageSize totally should be member of Pagination class otherwise this way looks so much functional way. Also it simplify unnecesary paramater pass in tiers of application. Second thing is that they use below interface. public interface IPagedList<T> : IList<T> { int PageCount { get; } int TotalItemCount { get; } int PageIndex { get; } int PageNumber { get; } int PageSize { get; } bool HasPreviousPage { get; } bool HasNextPage { get; } bool IsFirstPage { get; } bool IsLastPage { get; } } If i want to routing my pagination to different action so i have to create new view model for encapsulate action name in it or even controller name. Another solution can be that sending this interfaced model to view then specify action and controller hard coded in pager method as parameter but i am losing totally re-usability of my view because it is strictly depends on just one action. Another thing is that they use below code in view Html.Pager(Model.PageSize, Model.PageNumber, Model.TotalItemCount) If the model is IPagedList why they don't provide an overload method like @Html.Pager(Model) or even better one is @Html.Pager(). You know that we know model type in this way. Before i was doing mistake because i was using Model.PageIndex instead of Model.PageNumber. Another big issue is they strongly rely on IQueryable interface. How they know that i use IQueryable in my data layer ? I would expected that they work simply with collections that is keep pagination implementation persistence ignorant. What is wrong about my improvement ideas over their pagination implementations ? What is their reason to not implement their paginations in this way ?

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  • Turn off keyboard back-light Sony (VAIO SVF1521DCXW)

    - by KasiyA
    I have a Sony laptop and I want to turn keyboard back-light off. It doesn't have a shortcut function key for doing this on the keyboard . I can turn off it with VAIO Control Center in Windows but I don't know how can I turn it off in Ubuntu 14.04. There isn't available to me: /sys/devices/platform/sony-laptop/kbd_backlight doesn't exist on my machine. I have this folder /sys/devices/platform/sony-laptop/ and there is three folder one power folder and two shortcut-ed folder driver , subsystem and five file contains battery_care_health , battery_care_limiter , modalias , touchpad and event This is the output of running sudo modinfo sony-laptop: filename: /lib/modules/3.13.0-34-generic/kernel/drivers/platform/x86/sony-laptop.ko version: 0.6 license: GPL description: Sony laptop extras driver (SPIC and SNC ACPI device) author: Stelian Pop, Mattia Dongili srcversion: 5C6E050349475558A231C59 alias: acpi*:SNY6001:* alias: acpi*:SNY5001:* depends: intree: Y vermagic: 3.13.0-34-generic SMP mod_unload modversions signer: Magrathea: Glacier signing key sig_key: 50:0B:C5:C8:7D:4B:11:5C:F3:C1:50:4F:7A:92:E2:33:C6:14:3D:58 sig_hashalgo: sha512 parm: debug:set this to 1 (and RTFM) if you want to help the development of this driver (int) parm: no_spic:set this if you don't want to enable the SPIC device (int) parm: compat:set this if you want to enable backward compatibility mode (int) parm: mask:set this to the mask of event you want to enable (see doc) (ulong) parm: camera:set this to 1 to enable Motion Eye camera controls (only use it if you have a C1VE or C1VN model) (int) parm: minor:minor number of the misc device for the SPIC compatibility code, default is -1 (automatic) (int) parm: kbd_backlight:set this to 0 to disable keyboard backlight, 1 to enable it (default: no change from current value) (int) parm: kbd_backlight_timeout:meaningful values vary from 0 to 3 and their meaning depends on the model (default: no change from current value) (int) With the suggested command: sudo modprobe -r sony_laptop sudo modprobe -v sony_laptop kbd_backlight=0 Output was: insmod /lib/modules/3.13.0-34-generic/kernel/drivers/platform/x86/sony-laptop.ko kbd_backlight=0 It doesn't seem to affect the keyboard backlight. And also trying this command: sudo modprobe -v sony_laptop kbd_backlight_timeout=3 kbd_backlight=0 and doesn't seem to effect the keyboard backlight I also test it after restart laptop, And I didn't see any effect too. Important : By default, keyboard backlight is off; when I press a key it turns on and after 15 seconds it turns off again. It's the same result on battery and AC power I followed also http://ubuntuforums.org/showthread.php?t=2139597 and Keyboard backlighting not working on a Vaio VPCSB11FX but didn't work so.

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  • It's intellisense for SQL Server

    - by Nick Harrison
    It's intellisense for SQL Server Anyone who has ever worked with me, heard me speak, or read any of writings knows that I am a HUGE fan of Reflector.    By extension,  I am a big fan of Red - Gate   I have recently begun exploring some of their other offerings and came across this jewel. SQL Prompt is a plug in for Visual Studio and SQL Server Management Studio.    It provides several tools to make dealing with SQL a little easier for your friendly neighborhood developer. When you a query window in a database, the plugin kicks in and gathers the metadata for the database that you are in.    As you type a query, you get handy feedback like a list of tables after you type select.    You can select one of the tables, specify * and then tab to expand the select clause to include all of the columns from the selected table.    As you are building up the where clause, you are prompted by the names of columns in the selected tables. If you spend any time writing ad hoc queries or building stored procedures by hand, this can save you substantial time. If you are learning a new data model, this can greatly cut down on your frustration level. The other really cool thing here is Format SQL.   I have searched all over the place for a really good SQL formatter.    Badly formatted  SQL is so much harder to read than well formatted SQL.   Unfortunately, management studio offers no support for keeping your SQL well formatted.    There are many tools available to format your SQL.   Some work better than others.    Some don't work that well at all.   Most will give you some measure of control over how the formatted SQL looks.    SQL Prompt produces good results and is easy to configure. Sadly no tool is perfect, and what would we be without a wish list.    There are some features that I would like to see: Make it easier to paste SQL in and out of code.    Strip off string builder, etc Automate replacing hard coded values with bind variables or parameters In addition to reformatting SQL, which is a huge refactor, support for other SQL refactors would be nice.    Convert join to sub query and vice versa come to mind Wish list a side, this is a wonderful tool that easily saves me an hour or more on most weeks.

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  • Tips on ensuring Model Quality

    - by [email protected]
    Given enough data that represents well the domain and models that reflect exactly the decision being optimized, models usually provide good predictions that ensure lift. Nevertheless, sometimes the modeling situation is less than ideal. In this blog entry we explore the problems found in a few such situations and how to avoid them.1 - The Model does not reflect the problem you are trying to solveFor example, you may be trying to solve the problem: "What product should I recommend to this customer" but your model learns on the problem: "Given that a customer has acquired our products, what is the likelihood for each product". In this case the model you built may be too far of a proxy for the problem you are really trying to solve. What you could do in this case is try to build a model based on the result from recommendations of products to customers. If there is not enough data from actual recommendations, you could use a hybrid approach in which you would use the [bad] proxy model until the recommendation model converges.2 - Data is not predictive enoughIf the inputs are not correlated with the output then the models may be unable to provide good predictions. For example, if the input is the phase of the moon and the weather and the output is what car did the customer buy, there may be no correlations found. In this case you should see a low quality model.The solution in this case is to include more relevant inputs.3 - Not enough cases seenIf the data learned does not include enough cases, at least 200 positive examples for each output, then the quality of recommendations may be low. The obvious solution is to include more data records. If this is not possible, then it may be possible to build a model based on the characteristics of the output choices rather than the choices themselves. For example, instead of using products as output, use the product category, price and brand name, and then combine these models.4 - Output leaking into input giving the false impression of good quality modelsIf the input data in the training includes values that have changed or are available only because the output happened, then you will find some strong correlations between the input and the output, but these strong correlations do not reflect the data that you will have available at decision (prediction) time. For example, if you are building a model to predict whether a web site visitor will succeed in registering, and the input includes the variable DaysSinceRegistration, and you learn when this variable has already been set, you will probably see a big correlation between having a Zero (or one) in this variable and the fact that registration was successful.The solution is to remove these variables from the input or make sure they reflect the value as of the time of decision and not after the result is known. 

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  • Set up a TFS Server/Service demo environment in less than 1 minute now!

    - by Tarun Arora
    Release Notes – http://tfsdemosetup.codeplex.com/  | Download | Source Code | Report a Bug | Ideas To Demonstrate the capabilities of TFS 2012 Server/Service Task board you would need to set up TFS with some teams, a few team members, some sample stories, tasks, etc. That’s too many steps if you as me! Hi! My name is Tarun Arora, I am a Microsoft MVP in Visual Studio ALM & a Visual Studio ALM Ranger, as a consultant I have had to demo TFS Preview to potential customers several times a day. I usually create the team project during the demo to show off how quick and efficient it is, but setting up teams, team members, tasks usually takes longer I don’t prefer carrying out these steps during the demo. I have developed a .net based console application which uses the TFS API to create a standard demo environment saving me from all these manual steps. The console application reads the set up information from an XML file, leaving the setup process highly customizable. Figure 1 – Demo Dictionary, change values here for unique setup The console application today sets up, 1. Create a new Team 2. Set the team as the default team 3. Configure team settings      a. Set Backlog Iteration path      b. Set Team Iterations and start & finish dates      c. Set Team Area path 4. Add Team Members 5. Add Product Backlog Items & linked Tasks. Image 2 – The team website before (on the left) and after (on the right) running the console app Image 3 – Team configuration before (on left) and after (on right) with new team Demo and 2 members Image 4 – Iteration configuration before (on left) and after (on right) with new backlog iteration path & sprint dates set Image 5 – Area configuration (on left) and after (on right) with area path configured for the team   Image 6 – A demo ready Task Board and Task Board for Team Members Credits, - Mattias Sköld [Visual Studio ALM Ranger] – I have used TfsTeamTools to perform team creation & add members - Ivan Popek – TFS 2012 API blog posts had some fantastic reusable samples.  - Shai Raiten [Microsoft ALM MVP] – Great collection of posts on TFS API. Enjoy!

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  • Beware of const members

    - by nmarun
    I happened to learn a new thing about const today and how one needs to be careful with its usage. Let’s say I have a third-party assembly ‘ConstVsReadonlyLib’ with a class named ConstSideEffect.cs: 1: public class ConstSideEffect 2: { 3: public static readonly int StartValue = 10; 4: public const int EndValue = 20; 5: } In my project, I reference the above assembly as follows: 1: static void Main(string[] args) 2: { 3: for (int i = ConstSideEffect.StartValue; i < ConstSideEffect.EndValue; i++) 4: { 5: Console.WriteLine(i); 6: } 7: Console.ReadLine(); 8: } You’ll see values 10 through 19 as expected. Now, let’s say I receive a new version of the ConstVsReadonlyLib. 1: public class ConstSideEffect 2: { 3: public static readonly int StartValue = 5; 4: public const int EndValue = 30; 5: } If I just drop this new assembly in the bin folder and run the application, without rebuilding my console application, my thinking was that the output would be from 5 to 29. Of course I was wrong… if not you’d not be reading this blog. The actual output is from 5 through 19. The reason is due to the behavior of const and readonly members. To begin with, const is the compile-time constant and readonly is a runtime constant. Next, when you compile the code, a compile-time constant member is replaced with the value of the constant in the code. But, the IL generated when you reference a read-only constant, references the readonly variable, not its value. So, the IL version of the Main method, after compilation actually looks something like: 1: static void Main(string[] args) 2: { 3: for (int i = ConstSideEffect.StartValue; i < 20; i++) 4: { 5: Console.WriteLine(i); 6: } 7: Console.ReadLine(); 8: } I’m no expert with this IL thingi, but when I look at the disassembled code of the exe file (using IL Disassembler), I see the following: I see our readonly member still being referenced by the variable name (ConstVsReadonlyLib.ConstSideEffect::StartValue) in line 0001. Then there’s the Console.WriteLine in line 000b and finally, see the value of 20 in line 0017. This, I’m pretty sure is our const member being replaced by its value which marks the upper bound of the ‘for’ loop. Now you know why the output was from 5 through 19. This definitely is a side-effect of having const members and one needs to be aware of it. While we’re here, I’d like to add a few other points about const and readonly members: const is slightly faster, but is less flexible readonly cannot be declared within a method scope const can be used only on primitive types (numbers and strings) Just wanted to share this before going to bed!

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  • Voxel Engine in Multiplayer?

    - by Oliver Schöning
    This is a question more out of Interest for now, because I am not even near to the point that I could create this project at the moment. I really like the progress on the Atomontage Engine. A Voxel engine that is WIP at the moment. I would like to create a Voxel SERVER eventually. First in JavaScript (That's what I am learning right now) later perhaps in C++ for speed. Remember, I am perfectly aware that this is very hard! This is a brainstorm for the next 10 years as for now. What I would like to achieve one day is a Multiplayer Game in the Browser where the voxels positions are updated by XYZ input from the server. The Browser Does only 3 things: sending player input to the server, updating Voxel positions send from the server and rendering the world. I imagine using something like the Three.js libary on the client side. So that would be my programming dream right there... Now to something simpler for the near future. Right now I am learning javascript. And I am making games with Construct2. (A really cool JavaScript "game maker") The plan is to create a 2D Voxel enviorment (Block Voxels) on the Socket.IO Server* and send the position of the Voxels and Players to the Client side which then positions the Voxel Blocks to the Server Output coordinates. I think that is a bit more manageable then the other bigger idea. And also there should be no worries about speed with this type of project in JavaScript (I hope). Extra Info: *I am using nodejs (Without really knowing what it does besides making Socket.IO work) So now some questions: Is the "dream project" doable in JavaScript? Or is C++ just the best option because it does not take as long to be interpreted at run time like JavaScript? What are the limitations? I can think of some: Need of a Powerful server depending on how much information the server has to process. Internet Speed; Sending the data of the Voxel positions to every player could add up being very high! The browser FPS might go down quickly if rendering to many objects. One way of fixing reducing the packages Could be to let the browser calculate some of the Voxel positions from Several Values. But that would slow down the Client side too. What about the more achievable project? I am almost 100% convinced that this is possible in JavaScript, and that there are several ways of doing this. This is just XY position Updating for now.. Hope this did make some sense. Please comment if you got something to say :D

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  • Wordpress Installation (on IIS and SQL Server)

    - by Davide Mauri
    To proceed with the installation of Wordpress on SQL Server and IIS, first of all, you need to do the following steps Create a database on SQL Server that will be used by Wordpress Create login that can access to the just created database and put the user into ddladmin, db_datareader, db_datawriter roles Download and unpack Wordpress 3.3.2 (latest version as of 27 May 2012) zip file into a directory of your choice Download the wp-db-abstraction 1.1.4 (latest version as of 27 May 2012) plugin from wordpress.org website Now that the basic action has been done, you can start to setup and configure your Wordpress installation. Unpack and follow the instructions in the README.TXT file to install the Database Abstraction Layer. Mainly you have to: Upload wp-db-abstraction.php and the wp-db-abstraction directory to wp-content/mu-plugins.  This should be parallel to your regular plugins directory.  If the mu-plugins directory does not exist, you must create it. Put the db.php file from inside the wp-db-abstraction.php directory to wp-content/db.php Now you can create an application pool in IIS like the following one Create a website, using the above Application Pool, that points to the folder where you unpacked Wordpress files. Be sure to give the “Write” permission to the IIS account, as pointed out in this (old, but still quite valid) installation manual: http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#iis Now you’re ready to go. Point your browser to the configured website and the Wordpress installation screen will be there for you. When you’re requested to enter information to connect to MySQL database, simply skip that page, leaving the default values. If you have installed the Database Abstraction Layer, another database installation screen will appear after the one used by MySQL, and here you can enter the configuration information needed to connect to SQL Server. After having finished the installation steps, you should be able to access and navigate your wordpress site.  A final touch, and it’s done: just add the needed rewrite rules http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#urlrewrite and that’s it! Well. Not really. Unfortunately the current (as of 27 May 2012) version of the Database Abstraction Layer (1.1.4) has some bugs. Luckily they can be quickly fixed: Backslash Fix http://wordpress.org/support/topic/plugin-wp-db-abstraction-fix-problems-with-backslash-usage Select Top 0 Fix Make the change to the file “.\wp-content\mu-plugins\wp-db-abstraction\translations\sqlsrv\translations.php” suggested by “debettap”   http://sourceforge.net/tracker/?func=detail&aid=3485384&group_id=315685&atid=1328061 And now you have a 100% working Wordpress installation on SQL Server! Since I also wanted to take advantage of SQL Server Full Text Search, I’ve created a very simple wordpress plugin to setup full-text search and to use it as website search engine: http://wpfts.codeplex.com/ Enjoy!

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  • Compute directional light frustum from view furstum points and light direction

    - by Fabian
    I'm working on a friends engine project and my task is to construct a new frustum from the light direction that overlaps the view frustum and possible shadow casters. The project already has a function that creates a frustum for this but its way to big and includes way to many casters (shadows) which can't be seen in the view frustum. Now the only parameter of this function are the normalized light direction vector and a view class which lets me extract the 8 view frustum points in world space. I don't have any additional infos about the scene. I have read some of the related Questions here but non seem to fit very well to my problem as they often just point to cascaded shadow maps. Sadly i can't use DX or openGl functions directly because this engine has a dedicated math library. From what i've read so far the steps are: Transform view frustum points into light space and find min/max x and y values (or sometimes minima and maxima of all three axis) and create a AABB using the min/max vectors. But what comes after this step? How do i transform this new AABB back to world space? What i've done so far: CVector3 Points[8], MinLight = CVector3(FLT_MAX), MaxLight = CVector3(FLT_MAX); for(int i = 0; i<8;++i){ Points[i] = Points[i] * WorldToShadowMapMatrix; MinLight = Math::Min(Points[i],MinLight); MaxLight = Math::Max(Points[i],MaxLight); } AABox box(MinLight,MaxLight); I don't think this is the right way to do it. The near plain probably has to extend into the direction of the light source to include potentional shadow casters. I've read the Microsoft article about cascaded shadow maps http://msdn.microsoft.com/en-us/library/windows/desktop/ee416307%28v=vs.85%29.aspx which also includes some sample code. But they seem to use the scenes AABB to determine the near and far plane which I can't since i cant access this information from the funtion I'm working in. Could you guys please link some example code which shows the calculation of such frustum? Thanks in advance! Additional questio: is there a way to construct a WorldToFrustum matrix that represents the above transformation?

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  • Multiple render targets and gamma correctness in Direct3D9

    - by Mario
    Let's say in a deferred renderer when building your G-Buffer you're going to render texture color, normals, depth and whatever else to your multiple render targets at once. Now if you want to have a gamma-correct rendering pipeline and you use regular sRGB textures as well as rendertargets, you'll need to apply some conversions along the way, because your filtering, sampling and calculations should happen in linear space, not sRGB space. Of course, you could store linear color in your textures and rendertargets, but this might very well introduce bad precision and banding issues. Reading from sRGB textures is easy: just set SRGBTexture = true; in your texture sampler in your HLSL effect code and the hardware does the conversion sRGB-linear for you. Writing to an sRGB rendertarget is theoretically easy, too: just set SRGBWriteEnable = true; in your effect pass in HLSL and your linear colors will be converted to sRGB space automatically. But how does this work with multiple rendertargets? I only want to do these corrections to the color textures and rendertarget, not to the normals, depth, specularity or whatever else I'll be rendering to my G-Buffer. Ok, so I just don't apply SRGBTexture = true; to my non-color textures, but when using SRGBWriteEnable = true; I'll do a gamma correction to all the values I write out to my rendertargets, no matter what I actually store there. I found some info on gamma over at Microsoft: http://msdn.microsoft.com/en-us/library/windows/desktop/bb173460%28v=vs.85%29.aspx For hardware that supports Multiple Render Targets (Direct3D 9) or Multiple-element Textures (Direct3D 9), only the first render target or element is written. If I understand correctly, SRGBWriteEnable should only be applied to the first rendertarget, but according to my tests it doesn't and is used for all rendertargets instead. Now the only alternative seems to be to handle these corrections manually in my shader and only correct the actual color output, but I'm not totally sure, that this'll not have any negative impact on color correctness. E.g. if the GPU does any blending or filtering or multisampling after the Linear-sRGB conversion... Do I even need gamma correction in this case, if I'm just writing texture color without lighting to my rendertarget? As far as I know, I DO need it because of the texture filtering and mip sampling happening in sRGB space instead, if I don't correct for it. Anyway, it'd be interesting to hear other people's solutions or thoughts about this.

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  • SQL SERVER – Concurrancy Problems and their Relationship with Isolation Level

    - by pinaldave
    Concurrency is simply put capability of the machine to support two or more transactions working with the same data at the same time. This usually comes up with data is being modified, as during the retrieval of the data this is not the issue. Most of the concurrency problems can be avoided by SQL Locks. There are four types of concurrency problems visible in the normal programming. 1)      Lost Update – This problem occurs when there are two transactions involved and both are unaware of each other. The transaction which occurs later overwrites the transactions created by the earlier update. 2)      Dirty Reads – This problem occurs when a transactions selects data that isn’t committed by another transaction leading to read the data which may not exists when transactions are over. Example: Transaction 1 changes the row. Transaction 2 changes the row. Transaction 1 rolls back the changes. Transaction 2 has selected the row which does not exist. 3)      Nonrepeatable Reads – This problem occurs when two SELECT statements of the same data results in different values because another transactions has updated the data between the two SELECT statements. Example: Transaction 1 selects a row, which is later on updated by Transaction 2. When Transaction A later on selects the row it gets different value. 4)      Phantom Reads – This problem occurs when UPDATE/DELETE is happening on one set of data and INSERT/UPDATE is happening on the same set of data leading inconsistent data in earlier transaction when both the transactions are over. Example: Transaction 1 is deleting 10 rows which are marked as deleting rows, during the same time Transaction 2 inserts row marked as deleted. When Transaction 1 is done deleting rows, there will be still rows marked to be deleted. When two or more transactions are updating the data, concurrency is the biggest issue. I commonly see people toying around with isolation level or locking hints (e.g. NOLOCK) etc, which can very well compromise your data integrity leading to much larger issue in future. Here is the quick mapping of the isolation level with concurrency problems: Isolation Dirty Reads Lost Update Nonrepeatable Reads Phantom Reads Read Uncommitted Yes Yes Yes Yes Read Committed No Yes Yes Yes Repeatable Read No No No Yes Snapshot No No No No Serializable No No No No I hope this 400 word small article gives some quick understanding on concurrency issues and their relation to isolation level. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Best practices for logging and tracing in .NET

    - by Levidad
    I've been reading a lot about tracing and logging, trying to find some golden rule for best practices in the matter, but there isn't any. People say that good programmers produce good tracing, but put it that way and it has to come from experience. I've also read similar questions in here and through the internet and they are not really the same thing I am asking or do not have a satisfying answer, maybe because the questions lack some detail. So, folks say that tracing should sort of replicate the experience of debugging the application in cases where you can't attach a debugger. It should provide enough context so that you can see which path is taken at each control point in the application. Going deeper, you can even distinguish between tracing and event logging, in that "event logging is different from tracing in that it captures major states rather than detailed flow of control". Now, say I want to do my tracing and logging using only the standard .NET classes, those in the System.Diagnostics namespace. I figured that the TraceSource class is better for the job than the static Trace class, because I want to differentiate among the trace levels and using the TraceSource class I can pass in a parameter informing the event type, while using the Trace class I must use Trace.WriteLineIf and then verify things like SourceSwitch.TraceInformation and SourceSwitch.TraceErrors, and it doesn't even have properties like TraceVerbose or TraceStart. With all that in mind, would you consider a good practice to do as follows: Trace a "Start" event when begining a method, which should represent a single logical operation or a pipeline, along with a string representation of the parameter values passed in to the method. Trace an "Information" event when inserting an item into the database. Trace an "Information" event when taking one path or another in an important if/else statement. Trace a "Critical" or "Error" in a catch block depending on weather this is a recoverable error. Trace a "Stop" event when finishing the execution of the method. And also, please clarify when best to trace Verbose and Warning event types. If you have examples of code with nice trace/logging and are willing to share, that would be excelent. Note: I've found some good information here, but still not what I am looking for: http://msdn.microsoft.com/en-us/magazine/ff714589.aspx Thanks in advance!

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  • SQL SERVER – Script to Update a Specific Column in Entire Database

    - by Pinal Dave
    Last week, I have received a very interesting question and I find in email and I really liked the question as I had to play around with SQL Script for a while to come up with the answer he was looking for. Please read the question and I believe that all of us face this kind of situation. “Pinal, In our database we have recently introduced ModifiedDate column in all of the tables. Now onwards any update happens in the row, we are updating current date and time to that field. Now here is the issue, when we added that field we did not update it with a default value because we were not sure when we will go live with the system so we let it be NULL. Now modification to the application went live yesterday and we are now updating this field. Here is where I need your help. We need to update all the tables in our database where we have column created ModifiedDate and now want to update with current datetime. As our system is already live since yesterday there are several thousands of the rows which are already updated with real world value so we do not want to update those values. Essentially, in our entire database where ever there is a ModifiedDate column and if it is NULL we want to update that with current date time?  Do you have a script for it?” Honestly I did not have such a script. This is very specific required but I was able to come up with two different methods how he can use this method. Method 1 : Using INFORMATION_SCHEMA SELECT 'UPDATE ' + T.TABLE_SCHEMA + '.' + T.TABLE_NAME + ' SET ModifiedDate = GETDATE() WHERE ModifiedDate IS NULL;' FROM INFORMATION_SCHEMA.TABLES T INNER JOIN INFORMATION_SCHEMA.COLUMNS C ON T.TABLE_NAME = C.TABLE_NAME AND c.COLUMN_NAME ='ModifiedDate' WHERE T.TABLE_TYPE = 'BASE TABLE' ORDER BY T.TABLE_SCHEMA, T.TABLE_NAME; Method 2: Using DMV SELECT 'UPDATE ' + SCHEMA_NAME(t.schema_id) + '.' + t.name + ' SET ModifiedDate = GETDATE() WHERE ModifiedDate IS NULL;' FROM sys.tables AS t INNER JOIN sys.columns c ON t.OBJECT_ID = c.OBJECT_ID WHERE c.name ='ModifiedDate' ORDER BY SCHEMA_NAME(t.schema_id), t.name; Above scripts will create an UPDATE script which will do the task which is asked. We can pretty much the update script to any other SELECT statement and retrieve any other data as well. Click to Download Scripts Reference: Pinal Dave (http://blog.sqlauthority.com)  Filed under: PostADay, SQL, SQL Authority, SQL Joins, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Informed TDD &ndash; Kata &ldquo;To Roman Numerals&rdquo;

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/05/28/informed-tdd-ndash-kata-ldquoto-roman-numeralsrdquo.aspxIn a comment on my article on what I call Informed TDD (ITDD) reader gustav asked how this approach would apply to the kata “To Roman Numerals”. And whether ITDD wasn´t a violation of TDD´s principle of leaving out “advanced topics like mocks”. I like to respond with this article to his questions. There´s more to say than fits into a commentary. Mocks and TDD I don´t see in how far TDD is avoiding or opposed to mocks. TDD and mocks are orthogonal. TDD is about pocess, mocks are about structure and costs. Maybe by moving forward in tiny red+green+refactor steps less need arises for mocks. But then… if the functionality you need to implement requires “expensive” resource access you can´t avoid using mocks. Because you don´t want to constantly run all your tests against the real resource. True, in ITDD mocks seem to be in almost inflationary use. That´s not what you usually see in TDD demonstrations. However, there´s a reason for that as I tried to explain. I don´t use mocks as proxies for “expensive” resource. Rather they are stand-ins for functionality not yet implemented. They allow me to get a test green on a high level of abstraction. That way I can move forward in a top-down fashion. But if you think of mocks as “advanced” or if you don´t want to use a tool like JustMock, then you don´t need to use mocks. You just need to stand the sight of red tests for a little longer ;-) Let me show you what I mean by that by doing a kata. ITDD for “To Roman Numerals” gustav asked for the kata “To Roman Numerals”. I won´t explain the requirements again. You can find descriptions and TDD demonstrations all over the internet, like this one from Corey Haines. Now here is, how I would do this kata differently. 1. Analyse A demonstration of TDD should never skip the analysis phase. It should be made explicit. The requirements should be formalized and acceptance test cases should be compiled. “Formalization” in this case to me means describing the API of the required functionality. “[D]esign a program to work with Roman numerals” like written in this “requirement document” is not enough to start software development. Coding should only begin, if the interface between the “system under development” and its context is clear. If this interface is not readily recognizable from the requirements, it has to be developed first. Exploration of interface alternatives might be in order. It might be necessary to show several interface mock-ups to the customer – even if that´s you fellow developer. Designing the interface is a task of it´s own. It should not be mixed with implementing the required functionality behind the interface. Unfortunately, though, this happens quite often in TDD demonstrations. TDD is used to explore the API and implement it at the same time. To me that´s a violation of the Single Responsibility Principle (SRP) which not only should hold for software functional units but also for tasks or activities. In the case of this kata the API fortunately is obvious. Just one function is needed: string ToRoman(int arabic). And it lives in a class ArabicRomanConversions. Now what about acceptance test cases? There are hardly any stated in the kata descriptions. Roman numerals are explained, but no specific test cases from the point of view of a customer. So I just “invent” some acceptance test cases by picking roman numerals from a wikipedia article. They are supposed to be just “typical examples” without special meaning. Given the acceptance test cases I then try to develop an understanding of the problem domain. I´ll spare you that. The domain is trivial and is explain in almost all kata descriptions. How roman numerals are built is not difficult to understand. What´s more difficult, though, might be to find an efficient solution to convert into them automatically. 2. Solve The usual TDD demonstration skips a solution finding phase. Like the interface exploration it´s mixed in with the implementation. But I don´t think this is how it should be done. I even think this is not how it really works for the people demonstrating TDD. They´re simplifying their true software development process because they want to show a streamlined TDD process. I doubt this is helping anybody. Before you code you better have a plan what to code. This does not mean you have to do “Big Design Up-Front”. It just means: Have a clear picture of the logical solution in your head before you start to build a physical solution (code). Evidently such a solution can only be as good as your understanding of the problem. If that´s limited your solution will be limited, too. Fortunately, in the case of this kata your understanding does not need to be limited. Thus the logical solution does not need to be limited or preliminary or tentative. That does not mean you need to know every line of code in advance. It just means you know the rough structure of your implementation beforehand. Because it should mirror the process described by the logical or conceptual solution. Here´s my solution approach: The arabic “encoding” of numbers represents them as an ordered set of powers of 10. Each digit is a factor to multiply a power of ten with. The “encoding” 123 is the short form for a set like this: {1*10^2, 2*10^1, 3*10^0}. And the number is the sum of the set members. The roman “encoding” is different. There is no base (like 10 for arabic numbers), there are just digits of different value, and they have to be written in descending order. The “encoding” XVI is short for [10, 5, 1]. And the number is still the sum of the members of this list. The roman “encoding” thus is simpler than the arabic. Each “digit” can be taken at face value. No multiplication with a base required. But what about IV which looks like a contradiction to the above rule? It is not – if you accept roman “digits” not to be limited to be single characters only. Usually I, V, X, L, C, D, M are viewed as “digits”, and IV, IX etc. are viewed as nuisances preventing a simple solution. All looks different, though, once IV, IX etc. are taken as “digits”. Then MCMLIV is just a sum: M+CM+L+IV which is 1000+900+50+4. Whereas before it would have been understood as M-C+M+L-I+V – which is more difficult because here some “digits” get subtracted. Here´s the list of roman “digits” with their values: {1, I}, {4, IV}, {5, V}, {9, IX}, {10, X}, {40, XL}, {50, L}, {90, XC}, {100, C}, {400, CD}, {500, D}, {900, CM}, {1000, M} Since I take IV, IX etc. as “digits” translating an arabic number becomes trivial. I just need to find the values of the roman “digits” making up the number, e.g. 1954 is made up of 1000, 900, 50, and 4. I call those “digits” factors. If I move from the highest factor (M=1000) to the lowest (I=1) then translation is a two phase process: Find all the factors Translate the factors found Compile the roman representation Translation is just a look-up. Finding, though, needs some calculation: Find the highest remaining factor fitting in the value Remember and subtract it from the value Repeat with remaining value and remaining factors Please note: This is just an algorithm. It´s not code, even though it might be close. Being so close to code in my solution approach is due to the triviality of the problem. In more realistic examples the conceptual solution would be on a higher level of abstraction. With this solution in hand I finally can do what TDD advocates: find and prioritize test cases. As I can see from the small process description above, there are two aspects to test: Test the translation Test the compilation Test finding the factors Testing the translation primarily means to check if the map of factors and digits is comprehensive. That´s simple, even though it might be tedious. Testing the compilation is trivial. Testing factor finding, though, is a tad more complicated. I can think of several steps: First check, if an arabic number equal to a factor is processed correctly (e.g. 1000=M). Then check if an arabic number consisting of two consecutive factors (e.g. 1900=[M,CM]) is processed correctly. Then check, if a number consisting of the same factor twice is processed correctly (e.g. 2000=[M,M]). Finally check, if an arabic number consisting of non-consecutive factors (e.g. 1400=[M,CD]) is processed correctly. I feel I can start an implementation now. If something becomes more complicated than expected I can slow down and repeat this process. 3. Implement First I write a test for the acceptance test cases. It´s red because there´s no implementation even of the API. That´s in conformance with “TDD lore”, I´d say: Next I implement the API: The acceptance test now is formally correct, but still red of course. This will not change even now that I zoom in. Because my goal is not to most quickly satisfy these tests, but to implement my solution in a stepwise manner. That I do by “faking” it: I just “assume” three functions to represent the transformation process of my solution: My hypothesis is that those three functions in conjunction produce correct results on the API-level. I just have to implement them correctly. That´s what I´m trying now – one by one. I start with a simple “detail function”: Translate(). And I start with all the test cases in the obvious equivalence partition: As you can see I dare to test a private method. Yes. That´s a white box test. But as you´ll see it won´t make my tests brittle. It serves a purpose right here and now: it lets me focus on getting one aspect of my solution right. Here´s the implementation to satisfy the test: It´s as simple as possible. Right how TDD wants me to do it: KISS. Now for the second equivalence partition: translating multiple factors. (It´a pattern: if you need to do something repeatedly separate the tests for doing it once and doing it multiple times.) In this partition I just need a single test case, I guess. Stepping up from a single translation to multiple translations is no rocket science: Usually I would have implemented the final code right away. Splitting it in two steps is just for “educational purposes” here. How small your implementation steps are is a matter of your programming competency. Some “see” the final code right away before their mental eye – others need to work their way towards it. Having two tests I find more important. Now for the next low hanging fruit: compilation. It´s even simpler than translation. A single test is enough, I guess. And normally I would not even have bothered to write that one, because the implementation is so simple. I don´t need to test .NET framework functionality. But again: if it serves the educational purpose… Finally the most complicated part of the solution: finding the factors. There are several equivalence partitions. But still I decide to write just a single test, since the structure of the test data is the same for all partitions: Again, I´m faking the implementation first: I focus on just the first test case. No looping yet. Faking lets me stay on a high level of abstraction. I can write down the implementation of the solution without bothering myself with details of how to actually accomplish the feat. That´s left for a drill down with a test of the fake function: There are two main equivalence partitions, I guess: either the first factor is appropriate or some next. The implementation seems easy. Both test cases are green. (Of course this only works on the premise that there´s always a matching factor. Which is the case since the smallest factor is 1.) And the first of the equivalence partitions on the higher level also is satisfied: Great, I can move on. Now for more than a single factor: Interestingly not just one test becomes green now, but all of them. Great! You might say, then I must have done not the simplest thing possible. And I would reply: I don´t care. I did the most obvious thing. But I also find this loop very simple. Even simpler than a recursion of which I had thought briefly during the problem solving phase. And by the way: Also the acceptance tests went green: Mission accomplished. At least functionality wise. Now I´ve to tidy up things a bit. TDD calls for refactoring. Not uch refactoring is needed, because I wrote the code in top-down fashion. I faked it until I made it. I endured red tests on higher levels while lower levels weren´t perfected yet. But this way I saved myself from refactoring tediousness. At the end, though, some refactoring is required. But maybe in a different way than you would expect. That´s why I rather call it “cleanup”. First I remove duplication. There are two places where factors are defined: in Translate() and in Find_factors(). So I factor the map out into a class constant. Which leads to a small conversion in Find_factors(): And now for the big cleanup: I remove all tests of private methods. They are scaffolding tests to me. They only have temporary value. They are brittle. Only acceptance tests need to remain. However, I carry over the single “digit” tests from Translate() to the acceptance test. I find them valuable to keep, since the other acceptance tests only exercise a subset of all roman “digits”. This then is my final test class: And this is the final production code: Test coverage as reported by NCrunch is 100%: Reflexion Is this the smallest possible code base for this kata? Sure not. You´ll find more concise solutions on the internet. But LOC are of relatively little concern – as long as I can understand the code quickly. So called “elegant” code, however, often is not easy to understand. The same goes for KISS code – especially if left unrefactored, as it is often the case. That´s why I progressed from requirements to final code the way I did. I first understood and solved the problem on a conceptual level. Then I implemented it top down according to my design. I also could have implemented it bottom-up, since I knew some bottom of the solution. That´s the leaves of the functional decomposition tree. Where things became fuzzy, since the design did not cover any more details as with Find_factors(), I repeated the process in the small, so to speak: fake some top level, endure red high level tests, while first solving a simpler problem. Using scaffolding tests (to be thrown away at the end) brought two advantages: Encapsulation of the implementation details was not compromised. Naturally private methods could stay private. I did not need to make them internal or public just to be able to test them. I was able to write focused tests for small aspects of the solution. No need to test everything through the solution root, the API. The bottom line thus for me is: Informed TDD produces cleaner code in a systematic way. It conforms to core principles of programming: Single Responsibility Principle and/or Separation of Concerns. Distinct roles in development – being a researcher, being an engineer, being a craftsman – are represented as different phases. First find what, what there is. Then devise a solution. Then code the solution, manifest the solution in code. Writing tests first is a good practice. But it should not be taken dogmatic. And above all it should not be overloaded with purposes. And finally: moving from top to bottom through a design produces refactored code right away. Clean code thus almost is inevitable – and not left to a refactoring step at the end which is skipped often for different reasons.   PS: Yes, I have done this kata several times. But that has only an impact on the time needed for phases 1 and 2. I won´t skip them because of that. And there are no shortcuts during implementation because of that.

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  • Finding the Twins when Implementing Catmull-Clark subdivision using Half-Edge mesh [migrated]

    - by Ailurus
    Note: The description became a little longer than expected. Do you know a readable implementation of this algorithm using this mesh? Please let me know! I'm trying to implement Catmull-Clark subdivision using Matlab (because later on the results have to be compared with some other stuff already implemented in Matlab). First try was with a Vertex-Face mesh, the algorithm works but it is of course not very efficient (since you need neighbouring information for edges and faces). Therefore, I'm now using a Half-Edge mesh (info), see also the paper of Lutz Kettner. Wikipedia link to the idea behind Catmull-Clark SDV: Wiki. My problem lies in finding the Twin HalfEdges, I'm just not sure how to do this. Below I'm describing my thoughts on the implementation, trying to keep it concise. Half-Edge mesh (using indices to Vertices/HalfEdges/Faces): Vertex (x,y,z,Outgoing_HalfEdge) HalfEdge (HeadVertex (or TailVertex, which one should I use), Next, Face, Twin). Face (HalfEdge) To keep it simple for now, assume that every face is a quadrilateral. The actual mesh is a list of Vertices, HalfEdges and Faces. The new mesh will consist of NewVertices, NewHalfEdges and NewFaces, like this (note: Number_... is the number of ...): NumberNewVertices: Number_Faces + Number_HalfEdges/2 + Number_Vertices NumberNewHalfEdges: 4 * 4 * NumberFaces NumberNewfaces: 4 * NumberFaces Catmull-Clark: Find the FacePoint (centroid) of each Face: --> Just average the x,y,z values of the vertices, save as a NewVertex. Find the EdgePoint of each HalfEdge: --> To prevent duplicates (each HalfEdge has a Twin which would result in the same HalfEdge) --> Only calculate EdgePoints of the HalfEdge which has the lowest index of the Pair. Update old Vertices Ok, now all the new Vertices are calculated (however, their Outgoing_HalfEdge is still unknown). Next step to save the new HalfEdges and Faces. This is the part causing me problems! Loop through each old Face, there are 4 new Faces to be created (because of the quadrilateral assumption) First create the 4 new HalfEdges per New Face, starting at the FacePoint to the Edgepoint Next a new HalfEdge from the EdgePoint to an Updated Vertex Another new one from the Updated Vertex to the next EdgePoint Finally the fourth new HalfEdge from the EdgePoint back to the FacePoint. The HeadVertex of each new HalfEdge is known, the Next HalfEdge too. The Face is also known (since it is the new face you're creating!). Only the Twin HalfEdge is unknown, how should I know this? By the way, while looping through the Vertices of the new Face, assign the Outgoing_HalfEdge to the Vertices. This is probably the place to find out which HalfEdge is the Twin. Finally, after the 4 new HalfEdges are created, save the Face with the HalfVertex index the last newly created HalfVertex. I hope this is clear, if needed I can post my (obviously not-yet-finished) Matlab code.

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  • Using Oracle BPM to Extend Oracle Applications

    - by Michelle Kimihira
    Author: Srikant Subramaniam, Senior Principal Product Manager, Oracle Fusion Middleware Customers often modify applications to meet their specific business needs - varying regulatory requirements, unique business processes, product mix transition, etc. Traditional implementation practices for such modifications are typically invasive in nature and introduce risk into projects, affect time-to-market and ease of use, and ultimately increase the costs of running and maintaining the applications. Another downside of these traditional implementation practices is that they literally cast the application in stone, making it difficult for end-users to tailor their individual work environments to meet specific needs, without getting IT involved. For many businesses, however, IT lacks the capacity to support such rapid business changes. As a result, adopting innovative solutions to change the economics of customization becomes an imperative rather than a choice. Let's look at a banking process in Siebel Financial Services and Oracle Policy Automation (OPA) using Oracle Business Process Management. This approach makes modifications simple, quick to implement and easy to maintain/upgrade. The process model is based on the Loan Origination Process Accelerator, i.e., a set of ready to deploy business solutions developed by Oracle using Business Process Management (BPM) 11g, containing customizable and extensible pre-built processes to fit specific customer requirements. This use case is a branch-based loan origination process. Origination includes a number of steps ranging from accepting a loan application, applicant identity and background verification (Know Your Customer), credit assessment, risk evaluation and the eventual disbursal of funds (or rejection of the application). We use BPM to model all of these individual tasks and integrate (via web services) with: Siebel Financial Services and (simulated) backend applications: FLEXCUBE for loan management, Background Verification and Credit Rating. The process flow starts in Siebel when a customer applies for loan, switches to OPA for eligibility verification and product recommendations, before handing it off to BPM for approvals. OPA Connector for Siebel simplifies integration with Siebel’s web services framework by saving directly into Siebel the results from the self-service interview. This combination of user input and product recommendation invokes the BPM process for loan origination. At the end of the approval process, we update Siebel and the financial app to complete the loop. We use BPM Process Spaces to display role-specific data via dashboards, including the ability to track the status of a given process (flow trace). Loan Underwriters have visibility into the product mix (loan categories), status of loan applications (count of approved/rejected/pending), volume and values of loans approved per processing center, processing times, requested vs. approved amount and other relevant business metrics. Summary Oracle recommends the use of Fusion Middleware as an extensions platform for applications. This approach makes modifications simple, quick to implement and easy to maintain/upgrade applications (by moving customizations away from applications to the process layer). It is also easier to manage processes that span multiple applications by using Oracle BPM. Additional Information Product Information on Oracle.com: Oracle Fusion Middleware Follow us on Twitter and Facebook Subscribe to our regular Fusion Middleware Newsletter

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  • Draw a never-ending line in XNA

    - by user2236165
    I am drawing a line in XNA which I want to never end. I also have a tool that moves forward in X-direction and a camera which is centered at this tool. However, when I reach the end of the viewport the lines are not drawn anymore. Here are some pictures to illustrate my problem: At the start the line goes across the whole screen, but as my tool moves forward, we reach the end of the line. Here are the method which draws the lines: private void DrawEvenlySpacedSprites (Texture2D texture, Vector2 point1, Vector2 point2, float increment) { var distance = Vector2.Distance (point1, point2); // the distance between two points var iterations = (int)(distance / increment); // how many sprites with be drawn var normalizedIncrement = 1.0f / iterations; // the Lerp method needs values between 0.0 and 1.0 var amount = 0.0f; if (iterations == 0) iterations = 1; for (int i = 0; i < iterations; i++) { var drawPoint = Vector2.Lerp (point1, point2, amount); spriteBatch.Draw (texture, drawPoint, Color.White); amount += normalizedIncrement; } } Here are the draw method in Game. The dots are my lines: protected override void Draw (GameTime gameTime) { graphics.GraphicsDevice.Clear(Color.Black); nyVector = nextVector (gammelVector); GraphicsDevice.SetRenderTarget (renderTarget); spriteBatch.Begin (); DrawEvenlySpacedSprites (dot, gammelVector, nyVector, 0.9F); spriteBatch.End (); GraphicsDevice.SetRenderTarget (null); spriteBatch.Begin (SpriteSortMode.Deferred, BlendState.AlphaBlend, null, null, null, null, camera.transform); spriteBatch.Draw (renderTarget, new Vector2 (), Color.White); spriteBatch.Draw (tool, new Vector2(toolPos.X - (tool.Width/2), toolPos.Y - (tool.Height/2)), Color.White); spriteBatch.End (); gammelVector = new Vector2 (nyVector.X, nyVector.Y); base.Draw (gameTime); } Here's the next vector-method, It just finds me a new point where the line should be drawn with a new X-coordinate between 100 and 200 pixels and a random Y-coordinate between the old vector Y-coordinate and the height of the viewport: Vector2 nextVector (Vector2 vector) { return new Vector2 (vector.X + r.Next(100, 200), r.Next ((int)(vector.Y - 100), viewport.Height)); } Can anyone point me in the right direction here? I'm guessing it has to do with the viewport.width, but I'm not quite sure how to solve it. Thank you for reading!

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