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  • Rails app deployment challenge, not finding database table in production.log

    - by Stefan M
    I'm trying to setup PasswordPusher as my first ruby app ever. Building and running the webrick server as instructed in README works fine. It was only when I tried to add Apache ProxyPass and ProxyPassReverse that the page load slowed down to several minutes. So I gave mod_passenger a whirl but now it's unable to find the password table. Here's what I get in log/production.log. Started GET "/" for 10.10.2.13 at Sun Jun 10 08:07:19 +0200 2012 Processing by PasswordsController#new as HTML Completed 500 Internal Server Error in 1ms ActiveRecord::StatementInvalid (Could not find table 'passwords'): app/controllers/passwords_controller.rb:77:in `new' app/controllers/passwords_controller.rb:77:in `new' While in log/private.log I get a lot more output so here's just a snippet but it looks to me like it's working with the database. Edit: This was actually old log output, maybe from db:create. Migrating to AddUserToPassword (20120220172426) (0.3ms) ALTER TABLE "passwords" ADD "user_id" integer (0.0ms) PRAGMA index_list("passwords") (0.2ms) CREATE INDEX "index_passwords_on_user_id" ON "passwords" ("user_id") (0.7ms) INSERT INTO "schema_migrations" ("version") VALUES ('20120220172426') (0.1ms) select sqlite_version(*) (0.1ms) SELECT "schema_migrations"."version" FROM "schema_migrations" (0.0ms) PRAGMA index_list("passwords") (0.0ms) PRAGMA index_info('index_passwords_on_user_id') (4.6ms) PRAGMA index_list("rails_admin_histories") (0.0ms) PRAGMA index_info('index_rails_admin_histories') (0.0ms) PRAGMA index_list("users") (4.8ms) PRAGMA index_info('index_users_on_unlock_token') (0.0ms) PRAGMA index_info('index_users_on_reset_password_token') (0.0ms) PRAGMA index_info('index_users_on_email') (0.0ms) PRAGMA index_list("views") In my vhost I have it set to use RailsEnv private. <VirtualHost *:80> # ProxyPreserveHost on # # ProxyPass / http://10.220.100.209:180/ # ProxyPassReverse / http://10.220.100.209:180/ DocumentRoot /var/www/pwpusher/public <Directory /var/www/pwpusher/public> allow from all Options -MultiViews </Directory> RailsEnv private ServerName pwpush.intranet ErrorLog /var/log/apache2/error.log LogLevel debug CustomLog /var/log/apache2/access.log combined </VirtualHost> My passenger.conf in mods-enabled is default for Debian. <IfModule mod_passenger.c> PassengerRoot /usr PassengerRuby /usr/bin/ruby </IfModule> In the apache error.log I get something more cryptic to me. [Sun Jun 10 06:25:07 2012] [notice] Apache/2.2.16 (Debian) Phusion_Passenger/2.2.11 PHP/5.3.3-7+squeeze9 with Suhosin-Patch mod_ssl/2.2.16 OpenSSL/0.9.8o configured -- resuming normal operations /var/www/pwpusher/vendor/bundle/ruby/1.8/bundler/gems/modernizr-rails-09e9e6a92d67/lib/modernizr/rails/version.rb:3: warning: already initialized constant VERSION cache: [GET /] miss [Sun Jun 10 08:07:19 2012] [debug] mod_deflate.c(615): [client 10.10.2.13] Zlib: Compressed 728 to 423 : URL / /var/www/pwpusher/vendor/bundle/ruby/1.8/bundler/gems/modernizr-rails-09e9e6a92d67/lib/modernizr/rails/version.rb:3: warning: already initialized constant VERSION cache: [GET /] miss [Sun Jun 10 10:17:16 2012] [debug] mod_deflate.c(615): [client 10.10.2.13] Zlib: Compressed 728 to 423 : URL / Maybe that's routine stuff. I can see the rake command create files in the relative app root db/. I have private.sqlite3, production.sqlite3 among others. And here's my config/database.yml. base: &base adapter: sqlite3 timeout: 5000 development: database: db/development.sqlite3 <<: *base test: database: db/test.sqlite3 <<: *base private: database: db/private.sqlite3 <<: *base production: database: db/production.sqlite3 <<: *base I've tried setting absolute paths in it but that did not help.

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  • How to resolve high CPU + excessive stat("/etc/localtime") and clock_gettime(CLOCK_REALTIME) calls

    - by Yemster
    I've been experiencing really high CPU on a ruby on rails app (see stack below) and have been trying to diagnose the possible causes to no avail. Stack: ruby 1.9.3 rails 3.2.6 Apache/2.2.21 (Debian) Phusion Passenger 3.0.11 Whenever I run strace against the spiking Rack process PID (see Top excerpt below), I am seeing a tonne of stat("/etc/localtime") and clock_gettime(CLOCK_REALTIME) calls and have no idea how to stop these. Excerpt from Top showin running PID: PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ COMMAND 11674 www-user 20 0 313m 182m 5076 R 99 2.3 63:04.60 Rack: /var/www/my_rails_app/current 11634 www-user 20 0 411m 216m 5144 S 10 2.7 197:55.63 Rack: /var/www/my_rails_app/current Strace snippet below: [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 141474018}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 141577456}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 143073982}) = 0 [pid 11674] poll([{fd=15, events=POLLIN|POLLPRI}], 1, 0) = 0 (Timeout) [pid 11674] write(15, "b\0\0\0\3SELECT `images`.* FROM `ima"..., 102) = 102 [pid 11674] read(15, "\1\0\0\1\0229\0\0\2\3def\23myappy_productio"..., 16384) = 2063 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 144138035}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 ... [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] stat("/etc/localtime", {st_mode=S_IFREG|0644, st_size=118, ...}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 154076443}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 154189429}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 157185700}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 157298770}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 165076003}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 165212572}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 167542679}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354058955, 167683436}) = 0 .... [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 62052248}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 62182486}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 62919948}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 63057266}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 63751707}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 73730686}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 75874687}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 76077133}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 78205019}) = 0 ... [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 89370879}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 89583247}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 91637614}) = 0 [pid 11674] clock_gettime(CLOCK_REALTIME, {1354060036, 91782149}) = 0 Have Google'd around and came across a number of suggestions which I've tried with no success. Things tried so far: Have tried setting time zone as recommended here Made no difference and issue still persists. Content of my /etc/localtime: TZif2UTCTZif2UTC UTC0 Have tried the recommended fix for the leapsecond bug: date -s 'date' No joy so far. I'm fresh out of ideas so any help/advice on how to diagnose or resolve would be greatly appreciated.

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  • Apache URL rewriting in reverse proxy

    - by Jeremy Gooch
    I'm deploying Apache in front of a Karaf-hosted application (Apache and Karaf are on separate servers). I want Apache to operate as a reverse proxy and also to hide part of the URL. The URL to get the log-in page of the application directly from the app server is http://app-server:8181/jellyfish. Pages are served by the Jetty instance running within Karaf. Of course, this behaviour would usually be blocked by the firewall for everything except the reverse proxy server. With the firewall off, if you hit this URL then Jetty loads the log-in page. The browser's address bar correctly changes to http://app-server:8181/jellyfish/login?0 and everything works. What I want is for http://web-server (i.e. from the root) to map to Jetty on the app server with the name of the app (jellyfish) suppressed. e.g. The browser would change to show http://web-server/login?0 in the address bar and all subsequent URLs and content would be served with the web-server's domain and without the jellyfish clutter. I can get Apache to operate as a simple reverse proxy, using the following config (snippet):- ProxyPass /jellyfish http://app-server:8181/jellyfish ProxyPassReverse / http://app-server:8181/ ...but this requires the browser's URL to contain jellyfish and going to the root URL (http://web-server) gives a 404 Not Found. I've spent a lot of time trying to use mod_rewrite with and without its [P] flag to get around this, but without success. I then tried the ProxyPassMatch directive, but I can't seem to get that quite correct either. Here's the current config, as is loaded into /etc/apache2/sites-available/ on the web server. Note that there is a locally-hosted images directory. I've also kept the mod_rewrite proxy exploit protection and am suppressing a couple of mod_security rules that were giving false positives. <VirtualHost *:80> ServerAdmin admin@drummer-server ServerName drummer-server ErrorLog ${APACHE_LOG_DIR}/error.log LogLevel warn CustomLog ${APACHE_LOG_DIR}/access.log combined Alias /images/ "/var/www/images/" RewriteEngine On RewriteCond %{REQUEST_URI} !^$ RewriteCond %{REQUEST_URI} !^/ RewriteRule .* - [R=400,L] ProxyPass /images ! ProxyPassMatch ^/(.*) http://granny-server:8181/jellyfish/$1 ProxyPassReverse / http://granny-server:8181/jellyfish ProxyPreserveHost On SecRuleRemoveById 981059 981060 <Directory "/var/www/images"> Options Indexes MultiViews FollowSymLinks AllowOverride None Order allow,deny Allow from all </Directory> </VirtualHost> If I go to http://web-server, I get redirected to http://web-server/jellyfish/home but this gives a 404, with a complaint about trying to access /jellyfish/jellyfish/home - NB the browser's address bar does not contain the double /jellyfish. HTTP ERROR 404 Problem accessing /jellyfish/jellyfish/home. Reason: Not Found And, if I go to http://web-server/login, I get redirected to http://web-server/jellyfish/login?0 but this gives a 404, with a complaint about trying to access /jellyfish/jellyfish/login. HTTP ERROR 404 Problem accessing /jellyfish/jellyfish/login. Reason: Not Found So, I'm guessing I'm somehow passing through the rules twice. I am also slightly bemused as to where the home bit of the URL comes from in the first example. Can someone point me in the right direction, please? Thanks, J.

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  • Mod_Rewrite w Apache mod_jrun22.so & ColdFusion 9 on cPanel

    - by Eddie B
    How can I utilize mod_rewrite at either the httpd.conf level or per-directory level when mod_jrun22 seems to have short-stopped the rewrite process for ColdFusion pages? I have a ColdFusion 9 based site running on Centos 5.8 w cPanel. cPanel uses EasyApache 3 to manage virtual host containers and as such the conf for mod_jrun22.so, /usr/local/apache/conf/includes/pre_main_global.conf, is loaded prior to the main httpd.conf with the domain specific rules for the container. My assertion is that .cfm pages are failing to be rewritten due to the mod_jk22.so module having priority in the directive chain. To note, I also have a WordPress blog in the site where the rewrites appear to be working fine. For example the following code to remove the index file works fine for php and fails with cfm ... .htaccess under /blog/ : This works Options -Indexes -Multiviews <IfModule mod_rewrite.c> RewriteEngine On RewriteBase /blog/ RewriteRule ^index\.php$ - [L] RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /blog/index.php [L] </IfModule> .htaccess under / : This does not work as expected. Apache serves the page. ASSERT: This would redirect to domain.com/ without index.cfm Options -Indexes -Multiviews <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteRule ^index\.cfm$ - [L] RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /index.cfm [L] </IfModule> .htaccess under / : This works I'm presuming this is working because the redirect is to another .cfm page and a 404 handler in Application.cfc ... Options -Indexes -Multiviews <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteRule ^.*\.cfm$ - [L] RewriteCond %{REQUEST_FILENAME} !-d RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{ENV:REDIRECT_STATUS} =404 RewriteRule . /404.cfm$ [L] </IfModule> I've attempted numerous different methods to rewrite .cfm urls ... Adding [PT], [L], [R], [NS], Moving the script to Directory blocks under httpd.conf --- all with the same results ... either the rewrite doesn't work or Apache crashes in an endless loop ... Any help would be greatly appreciated. Below is a single-visit rewrite log snippet for a request to /index.cfm ... the pass-through is taking effect before the rewrite ... cat rewrite_dump_mod | grep index.cfm [perdir /home/foo/public_html/] strip per-dir prefix: /home/foo/public_html/index.cfm -> index.cfm [perdir /home/foo/public_html/] applying pattern '^.*\.cfm$' to uri 'index.cfm' [perdir /home/foo/public_html/] pass through /home/foo/public_html/index.cfm [perdir /home/foo/public_html/] strip per-dir prefix: /home/foo/public_html/index.cfm -> index.cfm [perdir /home/foo/public_html/] applying pattern '^.*\.cfm$' to uri 'index.cfm' [perdir /home/foo/public_html/] pass through /home/foo/public_html/index.cfm * UPDATE * I've managed to figure this out ... it took a while ... Options -Indexes -Multiviews +FollowSymLinks <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteCond %{HTTP_HOST} !^www\. RewriteRule ^(.*)$ http://www.%{HTTP_HOST}/$1 [R=301,L] RewriteCond %{THE_REQUEST} ^.*/index\.cfm RewriteRule ^(.*)index.cfm http://%{HTTP_HOST}/$1 [R=301,L] </IfModule>

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  • Twitter traffic might not be what it seems

    - by Piet
    Are you using bit.ly stats to measure interest in the links you post on twitter? I’ve been hearing for a while about people claiming to get the majority of their traffic originating from twitter these days. Now, I’ve been playing with the twitter ruby gem recently, doing various experiments which I’ll not go into detail here because they could be regarded as spamming… if I’d conduct them on a large scale, that is. It’s scary to see people actually engaging with @replies crafted with some regular expressions and eliza-like trickery on status updates found using the twitter api. I’m wondering how Twitter is going to contain the coming spam-flood. When posting links I used bit.ly as url shortener, since this one seems to be the de-facto standard on twitter. A nice thing about bit.ly is that it shows some basic stats about the redirects it performs for your shortened links. To my surprise, most links posted almost immediately resulted in several visitors. Now, seeing that I was posting the links together with some information concerning what the link is about, I concluded that the people who were actually clicking the links should be very targeted visitors. This felt a bit like free adwords, and I suddenly started to understand why everyone was raving about getting traffic from twitter. How wrong I was! (and I think several 1000 online marketers with me) On the destination site I used a traffic logging solution that works by including a little javascript snippet in your pages. It seemed that somehow all visitors disappeared after the bit.ly redirect and before getting to the site, because I was hardly seeing any visitors there. So I started investigating what was happening: by looking at the logfiles of the destination site, and by making my own ’shortened’ urls by doing redirects using a very short domain name I own. This way, I could check the apache access_log before the redirects. Most user agents turned out to be bots without a doubt. Here’s an excerpt of user-agents awk’ed from apache’s access_log for a time period of about one hour, right after posting some links: AideRSS 2.0 (postrank.com) Java/1.6.0_13 Java/1.6.0_14 libwww-perl/5.816 MLBot (www.metadatalabs.com/mlbot) Mozilla/4.0 (compatible;MSIE 5.01; Windows -NT 5.0 - real-url.org) Mozilla/5.0 (compatible; Twitturls; +http://twitturls.com) Mozilla/5.0 (compatible; Viralheat Bot/1.0; +http://www.viralheat.com/) Mozilla/5.0 (Danger hiptop 4.6; U; rv:1.7.12) Gecko/20050920 Mozilla/5.0 (X11; U; Linux i686; en-us; rv:1.9.0.2) Gecko/2008092313 Ubuntu/9.04 (jaunty) Firefox/3.5 OpenCalaisSemanticProxy PycURL/7.18.2 PycURL/7.19.3 Python-urllib/1.17 Twingly Recon twitmatic Twitturly / v0.6 Wget/1.10.2 (Red Hat modified) Wget/1.11.1 (Red Hat modified) Of the few user-agents that seem ‘real’ at first, half are originating from an ip-address used by Amazon EC2. And I doubt people are setting op proxies on there. Oh yeah, Googlebot (the real deal, from a legit google owned address) is sucking up posted links like fresh oysters. I guess google is trying to make sure in advance to never be beaten by twitter in the ‘realtime search’ department. Actually, I think it’d be almost stupid NOT to post any new pages/posts/websites on Twitter, it must be one of the fastest ways to get a Googlebot visit. Same experiment with a real, established twitter account Now, because I was posting the url’s either as ’status’ messages or directed @people, on a test-account with hardly any (human) followers, I checked again using the twitter accounts from a commercial site I’m involved with. These accounts all have between 500 and 1000 targeted (I think) followers. I checked the destination access_logs and also added ‘my’ redirect after the bit.ly redirect: same results, although seemingly a bit higher real visitor/bot ratio. Btw: one of these account was ‘punished’ with a 1 week lock recently because the same (1 one!) status update was sent that was sent right before using another account. They got an email explaining the lock because the account didn’t act according to their TOS. I can’t find anything in their TOS about it, can you? I don’t think Twitter is on the right track punishing a legit account, knowing the trickery I had been doing with it’s api went totally unpunished. I might be wrong though, I often am. On the other hand: this commercial site reported targeted traffic and actual signups from visitors coming from Twitter. The ones that are really real visitors are also very targeted. I’m just not sure if the amount of work involved could hold up against an adwords campaign. Reposting the same link over and over again helps On thing I noticed: It helps to keep on reposting the same links with regular intervals. I guess most people only look at their first page when checking out recent posts of the ones they’re following, or don’t look too far back when performing a search. Now, this probably isn’t according to the twitter TOS. Actually, it might be spamming but no-one is obligated to follow anyone else of course. This way, I was getting more real visitors and less bots. To my surprise (when my programmer’s hat is on) there were still repeated visits from the same bots coming from the same ip-addresses. Did they expect to find something else when visiting for a 2nd or 3rd time? (actually,this gave me an idea: you can’t change a link once it’s posted, but you can change where it redirects to) Most bots were smart enough not to follow the same link again though. Are you successful in getting real visitors from Twitter? Are you only relying on bit.ly to provide traffic stats?

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  • BNF – how to read syntax?

    - by Piotr Rodak
    A few days ago I read post of Jen McCown (blog) about her idea of blogging about random articles from Books Online. I think this is a great idea, even if Jen says that it’s not exciting or sexy. I noticed that many of the questions that appear on forums and other media arise from pure fact that people asking questions didn’t bother to read and understand the manual – Books Online. Jen came up with a brilliant, concise acronym that describes very well the category of posts about Books Online – RTFM365. I take liberty of tagging this post with the same acronym. I often come across questions of type – ‘Hey, i am trying to create a table, but I am getting an error’. The error often says that the syntax is invalid. 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT DEFAULT Guid_Default NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); 5 The answer is usually(1), ‘Ok, let me check it out.. Ah yes – you have to put name of the DEFAULT constraint before the type of constraint: 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT Guid_Default DEFAULT NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); Why many people stumble on syntax errors? Is the syntax poorly documented? No, the issue is, that correct syntax of the CREATE TABLE statement is documented very well in Books Online and is.. intimidating. Many people can be taken aback by the rather complex block of code that describes all intricacies of the statement. However, I don’t know better way of defining syntax of the statement or command. The notation that is used to describe syntax in Books Online is a form of Backus-Naur notatiion, called BNF for short sometimes. This is a notation that was invented around 50 years ago, and some say that even earlier, around 400 BC – would you believe? Originally it was used to define syntax of, rather ancient now, ALGOL programming language (in 1950’s, not in ancient India). If you look closer at the definition of the BNF, it turns out that the principles of this syntax are pretty simple. Here are a few bullet points: italic_text is a placeholder for your identifier <italic_text_in_angle_brackets> is a definition which is described further. [everything in square brackets] is optional {everything in curly brackets} is obligatory everything | separated | by | operator is an alternative ::= “assigns” definition to an identifier Yes, it looks like these six simple points give you the key to understand even the most complicated syntax definitions in Books Online. Books Online contain an article about syntax conventions – have you ever read it? Let’s have a look at fragment of the CREATE TABLE statement: 1 CREATE TABLE 2 [ database_name . [ schema_name ] . | schema_name . ] table_name 3 ( { <column_definition> | <computed_column_definition> 4 | <column_set_definition> } 5 [ <table_constraint> ] [ ,...n ] ) 6 [ ON { partition_scheme_name ( partition_column_name ) | filegroup 7 | "default" } ] 8 [ { TEXTIMAGE_ON { filegroup | "default" } ] 9 [ FILESTREAM_ON { partition_scheme_name | filegroup 10 | "default" } ] 11 [ WITH ( <table_option> [ ,...n ] ) ] 12 [ ; ] Let’s look at line 2 of the above snippet: This line uses rules 3 and 5 from the list. So you know that you can create table which has specified one of the following. just name – table will be created in default user schema schema name and table name – table will be created in specified schema database name, schema name and table name – table will be created in specified database, in specified schema database name, .., table name – table will be created in specified database, in default schema of the user. Note that this single line of the notation describes each of the naming schemes in deterministic way. The ‘optionality’ of the schema_name element is nested within database_name.. section. You can use either database_name and optional schema name, or just schema name – this is specified by the pipe character ‘|’. The error that user gets with execution of the first script fragment in this post is as follows: Msg 156, Level 15, State 1, Line 2 Incorrect syntax near the keyword 'DEFAULT'. Ok, let’s have a look how to find out the correct syntax. Line number 3 of the BNF fragment above contains reference to <column_definition>. Since column_definition is in angle brackets, we know that this is a reference to notion described further in the code. And indeed, the very next fragment of BNF contains syntax of the column definition. 1 <column_definition> ::= 2 column_name <data_type> 3 [ FILESTREAM ] 4 [ COLLATE collation_name ] 5 [ NULL | NOT NULL ] 6 [ 7 [ CONSTRAINT constraint_name ] DEFAULT constant_expression ] 8 | [ IDENTITY [ ( seed ,increment ) ] [ NOT FOR REPLICATION ] 9 ] 10 [ ROWGUIDCOL ] [ <column_constraint> [ ...n ] ] 11 [ SPARSE ] Look at line 7 in the above fragment. It says, that the column can have a DEFAULT constraint which, if you want to name it, has to be prepended with [CONSTRAINT constraint_name] sequence. The name of the constraint is optional, but I strongly recommend you to make the effort of coming up with some meaningful name yourself. So the correct syntax of the CREATE TABLE statement from the beginning of the article is like this: 1 CREATE TABLE dbo.Employees 2 (guid uniqueidentifier CONSTRAINT Guid_Default DEFAULT NEWSEQUENTIALID() ROWGUIDCOL, 3 Employee_Name varchar(60) 4 CONSTRAINT Guid_PK PRIMARY KEY (guid) ); That is practically everything you should know about BNF. I encourage you to study the syntax definitions for various statements and commands in Books Online, you can find really interesting things hidden there. Technorati Tags: SQL Server,t-sql,BNF,syntax   (1) No, my answer usually is a question – ‘What error message? What does it say?’. You’d be surprised to know how many people think I can go through time and space and look at their screen at the moment they received the error.

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  • Visual Studio 2013, ASP.NET MVC 5 Scaffolded Controls, and Bootstrap

    - by plitwin
    A few days ago, I created an ASP.NET MVC 5 project in the brand new Visual Studio 2013. I added some model classes and then proceeded to scaffold a controller class and views using the Entity Framework. Scaffolding Some Views Visual Studio 2013, by default, uses the Bootstrap 3 responsive CSS framework. Great; after all, we all want our web sites to be responsive and work well on mobile devices. Here’s an example of a scaffolded Create view as shown in Google Chrome browser   Looks pretty good. Okay, so let’s increase the width of the Title, Description, Address, and Date/Time textboxes. And decrease the width of the  State and MaxActors textbox controls. Can’t be that hard… Digging Into the Code Let’s take a look at the scaffolded Create.cshtml file. Here’s a snippet of code behind the Create view. Pretty simple stuff. @using (Html.BeginForm()) { @Html.AntiForgeryToken() <div class="form-horizontal"> <h4>RandomAct</h4> <hr /> @Html.ValidationSummary(true) <div class="form-group"> @Html.LabelFor(model => model.Title, new { @class = "control-label col-md-2" }) <div class="col-md-10"> @Html.EditorFor(model => model.Title) @Html.ValidationMessageFor(model => model.Title) </div> </div> <div class="form-group"> @Html.LabelFor(model => model.Description, new { @class = "control-label col-md-2" }) <div class="col-md-10"> @Html.EditorFor(model => model.Description) @Html.ValidationMessageFor(model => model.Description) </div> </div> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } A little more digging and I discover that there are three CSS files of importance in how the page is rendered: boostrap.css (and its minimized cohort) and site.css as shown below.   The Root of the Problem And here’s the root of the problem which you’ll find the following CSS in Site.css: /* Set width on the form input elements since they're 100% wide by default */ input, select, textarea { max-width: 280px; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Yes, Microsoft is for some reason setting the maximum width of all input, select, and textarea controls to 280 pixels. Not sure the motivation behind this, but until you change this or overrride this by assigning the form controls to some other CSS class, your controls will never be able to be wider than 280px. The Fix Okay, so here’s the deal: I hope to become very competent in all things Bootstrap in the near future, but I don’t think you should have to become a Bootstrap guru in order to modify some scaffolded control widths. And you don’t. Here is the solution I came up with: Find the aforementioned CSS code in SIte.css and change it to something more tenable. Such as: /* Set width on the form input elements since they're 100% wide by default */ input, select, textarea { max-width: 600px; } Because the @Html.EditorFor html helper doesn’t support the passing of HTML attributes, you will need to repalce any @Html.EditorFor() helpers with @Html.TextboxFor(), @Html.TextAreaFor, @Html.CheckBoxFor, etc. helpers, and then add a custom width attribute to each control you wish to modify. Thus, the earlier stretch of code might end up looking like this: @using (Html.BeginForm()) { @Html.AntiForgeryToken() <div class="form-horizontal"> <h4>Random Act</h4> <hr /> @Html.ValidationSummary(true) <div class="form-group"> @Html.LabelFor(model => model.Title, new { @class = "control-label col-md-2" }) <div class="col-md-10"> @Html.TextBoxFor(model => model.Title, new { style = "width: 400px" }) @Html.ValidationMessageFor(model => model.Title) </div> </div> <div class="form-group"> @Html.LabelFor(model => model.Description, new { @class = "control-label col-md-2" }) <div class="col-md-10"> @Html.TextAreaFor(model => model.Description, new { style = "width: 400px" }) @Html.ValidationMessageFor(model => model.Description) </div> </div> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Resulting Form Here’s what the page looks like after the fix: Technorati Tags: ASP.NET MVC,ASP.NET MVC 5,Bootstrap

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  • Overwriting TFS Web Services

    - by javarg
    In this blog I will share a technique I used to intercept TFS Web Services calls. This technique is a very invasive one and requires you to overwrite default TFS Web Services behavior. I only recommend taking such an approach when other means of TFS extensibility fail to provide the same functionality (this is not a supported TFS extensibility point). For instance, intercepting and aborting a Work Item change operation could be implemented using this approach (consider TFS Subscribers functionality before taking this approach, check Martin’s post about subscribers). So let’s get started. The technique consists in versioning TFS Web Services .asmx service classes. If you look into TFS’s ASMX services you will notice that versioning is supported by creating a class hierarchy between different product versions. For instance, let’s take the Work Item management service .asmx. Check the following .asmx file located at: %Program Files%\Microsoft Team Foundation Server 2010\Application Tier\Web Services\_tfs_resources\WorkItemTracking\v3.0\ClientService.asmx The .asmx references the class Microsoft.TeamFoundation.WorkItemTracking.Server.ClientService3: <%-- Copyright (c) Microsoft Corporation. All rights reserved. --%> <%@ webservice language="C#" Class="Microsoft.TeamFoundation.WorkItemTracking.Server.ClientService3" %> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The inheritance hierarchy for this service class follows: Note the naming convention used for service versioning (ClientService3, ClientService2, ClientService). We will need to overwrite the latest service version provided by the product (in this case ClientService3 for TFS 2010). The following example intercepts and analyzes WorkItem fields. Suppose we need to validate state changes with more advanced logic other than the provided validations/constraints of the process template. Important: Backup the original .asmx file and create one of your own. Create a Visual Studio Web App Project and include a new ASMX Web Service in the project Add the following references to the project (check the folder %Program Files%\Microsoft Team Foundation Server 2010\Application Tier\Web Services\bin\): Microsoft.TeamFoundation.Framework.Server.dll Microsoft.TeamFoundation.Server.dll Microsoft.TeamFoundation.Server.dll Microsoft.TeamFoundation.WorkItemTracking.Client.QueryLanguage.dll Microsoft.TeamFoundation.WorkItemTracking.Server.DataAccessLayer.dll Microsoft.TeamFoundation.WorkItemTracking.Server.DataServices.dll Replace the default service implementation with the something similar to the following code: Code Snippet /// <summary> /// Inherit from ClientService3 to overwrite default Implementation /// </summary> [WebService(Namespace = "http://schemas.microsoft.com/TeamFoundation/2005/06/WorkItemTracking/ClientServices/03", Description = "Custom Team Foundation WorkItemTracking ClientService Web Service")] public class CustomTfsClientService : ClientService3 {     [WebMethod, SoapHeader("requestHeader", Direction = SoapHeaderDirection.In)]     public override bool BulkUpdate(         XmlElement package,         out XmlElement result,         MetadataTableHaveEntry[] metadataHave,         out string dbStamp,         out Payload metadata)     {         var xe = XElement.Parse(package.OuterXml);         // We only intercept WorkItems Updates (we can easily extend this sample to capture any operation).         var wit = xe.Element("UpdateWorkItem");         if (wit != null)         {             if (wit.Attribute("WorkItemID") != null)             {                 int witId = (int)wit.Attribute("WorkItemID");                 // With this Id. I can query TFS for more detailed information, using TFS Client API (assuming the WIT already exists).                 var stateChanged =                     wit.Element("Columns").Elements("Column").FirstOrDefault(c => (string)c.Attribute("Column") == "System.State");                 if (stateChanged != null)                 {                     var newStateName = stateChanged.Element("Value").Value;                     if (newStateName == "Resolved")                     {                         throw new Exception("Cannot change state to Resolved!");                     }                 }             }         }         // Finally, we call base method implementation         return base.BulkUpdate(package, out result, metadataHave, out dbStamp, out metadata);     } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } 4. Build your solution and overwrite the original .asmx with the new implementation referencing our new service version (don’t forget to backup it up first). 5. Copy your project’s .dll into the following path: %Program Files%\Microsoft Team Foundation Server 2010\Application Tier\Web Services\bin 6. Try saving a WorkItem into the Resolved state. Enjoy!

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  • Routing Issue in ASP.NET MVC 3 RC 2

    - by imran_ku07
         Introduction:             Two weeks ago, ASP.NET MVC team shipped the ASP.NET MVC 3 RC 2 release. This release includes some new features and some performance optimization. This release also fixes most of the bugs but still some minor issues are present in this release. Some of these issues are already discussed by Scott Guthrie at Update on ASP.NET MVC 3 RC2 (and a workaround for a bug in it). In addition to these issues, I have found another issue in this release regarding routing. In this article, I will show you the issue regarding routing and a simple workaround for this issue.       Description:             The easiest way to understand an issue is to reproduce it in the application. So create a MVC 2 application and a MVC 3 RC 2 application. Then in both applications, just open global.asax file and update the default route as below,     routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); routes.MapRoute( "Default", // Route name "{controller}/{action}/{id1}/{id2}", // URL with parameters new { controller = "Home", action = "Index", id1 = UrlParameter.Optional, id2 = UrlParameter.Optional } // Parameter defaults );              Then just open Index View and add the following lines,    <%@ Page Language="C#" MasterPageFile="~/Views/Shared/Site.Master" Inherits="System.Web.Mvc.ViewPage" %> <asp:Content ID="Content1" ContentPlaceHolderID="TitleContent" runat="server"> Home Page </asp:Content> <asp:Content ID="Content2" ContentPlaceHolderID="MainContent" runat="server"> <% Html.RenderAction("About"); %> </asp:Content>             The above view will issue a child request to About action method. Now run both applications. ASP.NET MVC 2 application will run just fine. But ASP.NET MVC 3 RC 2 application will throw an exception as shown below,                  You may think that this is a routing issue but this is not the case here as both ASP.NET MVC 2 and ASP.NET MVC  3 RC 2 applications(created above) are built with .NET Framework 4.0 and both will use the same routing defined in System.Web. Something is wrong in ASP.NET MVC 3 RC 2. So after digging into ASP.NET MVC source code, I have found that the UrlParameter class in ASP.NET MVC 3 RC 2 overrides the ToString method which simply return an empty string.     public sealed class UrlParameter { public static readonly UrlParameter Optional = new UrlParameter(); private UrlParameter() { } public override string ToString() { return string.Empty; } }             In MVC 2 the ToString method was not overridden. So to quickly fix the above problem just replace UrlParameter.Optional default value with a different value other than null or empty(for example, a single white space) or replace UrlParameter.Optional default value with a new class object containing the same code as UrlParameter class have except the ToString method is not overridden (or with a overridden ToString method that return a string value other than null or empty). But by doing this you will loose the benefit of ASP.NET MVC 2 Optional URL Parameters. There may be many different ways to fix the above problem and not loose the benefit of optional parameters. Here I will create a new class MyUrlParameter with the same code as UrlParameter class have except the ToString method is not overridden. Then I will create a base controller class which contains a constructor to remove all MyUrlParameter route data parameters, same like ASP.NET MVC doing with UrlParameter route data parameters early in the request.     public class BaseController : Controller { public BaseController() { if (System.Web.HttpContext.Current.CurrentHandler is MvcHandler) { RouteValueDictionary rvd = ((MvcHandler)System.Web.HttpContext.Current.CurrentHandler).RequestContext.RouteData.Values; string[] matchingKeys = (from entry in rvd where entry.Value == MyUrlParameter.Optional select entry.Key).ToArray(); foreach (string key in matchingKeys) { rvd.Remove(key); } } } } public class HomeController : BaseController { public ActionResult Index(string id1) { ViewBag.Message = "Welcome to ASP.NET MVC!"; return View(); } public ActionResult About() { return Content("Child Request Contents"); } }     public sealed class MyUrlParameter { public static readonly MyUrlParameter Optional = new MyUrlParameter(); private MyUrlParameter() { } }     routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); routes.MapRoute( "Default", // Route name "{controller}/{action}/{id1}/{id2}", // URL with parameters new { controller = "Home", action = "Index", id1 = MyUrlParameter.Optional, id2 = MyUrlParameter.Optional } // Parameter defaults );             MyUrlParameter class is a copy of UrlParameter class except that MyUrlParameter class not overrides the ToString method. Note that the default route is modified to use MyUrlParameter.Optional instead of UrlParameter.Optional. Also note that BaseController class constructor is removing MyUrlParameter parameters from the current request route data so that the model binder will not bind these parameters with action method parameters. Now just run the ASP.NET MVC 3 RC 2 application again, you will find that it runs just fine.             In case if you are curious to know that why ASP.NET MVC 3 RC 2 application throws an exception if UrlParameter class contains a ToString method which returns an empty string, then you need to know something about a feature of routing for url generation. During url generation, routing will call the ParsedRoute.Bind method internally. This method includes a logic to match the route and build the url. During building the url, ParsedRoute.Bind method will call the ToString method of the route values(in our case this will call the UrlParameter.ToString method) and then append the returned value into url. This method includes a logic after appending the returned value into url that if two continuous returned values are empty then don't match the current route otherwise an incorrect url will be generated. Here is the snippet from ParsedRoute.Bind method which will prove this statement.       if ((builder2.Length > 0) && (builder2[builder2.Length - 1] == '/')) { return null; } builder2.Append("/"); ........................................................... ........................................................... ........................................................... ........................................................... if (RoutePartsEqual(obj3, obj4)) { builder2.Append(UrlEncode(Convert.ToString(obj3, CultureInfo.InvariantCulture))); continue; }             In the above example, both id1 and id2 parameters default values are set to UrlParameter object and UrlParameter class include a ToString method that returns an empty string. That's why this route will not matched.            Summary:             In this article I showed you the issue regarding routing and also showed you how to workaround this problem. I explained this issue with an example by creating a ASP.NET MVC 2 and a ASP.NET MVC 3 RC 2 application. Finally I also explained the reason for this issue. Hopefully you will enjoy this article too.   SyntaxHighlighter.all()

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  • Asserting with JustMock

    - by mehfuzh
    In this post, i will be digging in a bit deep on Mock.Assert. This is the continuation from previous post and covers up the ways you can use assert for your mock expectations. I have used another traditional sample of Talisker that has a warehouse [Collaborator] and an order class [SUT] that will call upon the warehouse to see the stock and fill it up with items. Our sample, interface of warehouse and order looks similar to : public interface IWarehouse {     bool HasInventory(string productName, int quantity);     void Remove(string productName, int quantity); }   public class Order {     public string ProductName { get; private set; }     public int Quantity { get; private set; }     public bool IsFilled { get; private set; }       public Order(string productName, int quantity)     {         this.ProductName = productName;         this.Quantity = quantity;     }       public void Fill(IWarehouse warehouse)     {         if (warehouse.HasInventory(ProductName, Quantity))         {             warehouse.Remove(ProductName, Quantity);             IsFilled = true;         }     }   }   Our first example deals with mock object assertion [my take] / assert all scenario. This will only act on the setups that has this “MustBeCalled” flag associated. To be more specific , let first consider the following test code:    var order = new Order(TALISKER, 0);    var wareHouse = Mock.Create<IWarehouse>();      Mock.Arrange(() => wareHouse.HasInventory(Arg.Any<string>(), 0)).Returns(true).MustBeCalled();    Mock.Arrange(() => wareHouse.Remove(Arg.Any<string>(), 0)).Throws(new InvalidOperationException()).MustBeCalled();    Mock.Arrange(() => wareHouse.Remove(Arg.Any<string>(), 100)).Throws(new InvalidOperationException());      //exercise    Assert.Throws<InvalidOperationException>(() => order.Fill(wareHouse));    // it will assert first and second setup.    Mock.Assert(wareHouse); Here, we have created the order object, created the mock of IWarehouse , then I setup our HasInventory and Remove calls of IWarehouse with my expected, which is called by the order.Fill internally. Now both of these setups are marked as “MustBeCalled”. There is one additional IWarehouse.Remove that is invalid and is not marked.   On line 9 ,  as we do order.Fill , the first and second setups will be invoked internally where the third one is left  un-invoked. Here, Mock.Assert will pass successfully as  both of the required ones are called as expected. But, if we marked the third one as must then it would fail with an  proper exception. Here, we can also see that I have used the same call for two different setups, this feature is called sequential mocking and will be covered later on. Moving forward, let’s say, we don’t want this must call, when we want to do it specifically with lamda. For that let’s consider the following code: //setup - data var order = new Order(TALISKER, 50); var wareHouse = Mock.Create<IWarehouse>();   Mock.Arrange(() => wareHouse.HasInventory(TALISKER, 50)).Returns(true);   //exercise order.Fill(wareHouse);   //verify state Assert.True(order.IsFilled); //verify interaction Mock.Assert(()=> wareHouse.HasInventory(TALISKER, 50));   Here, the snippet shows a case for successful order, i haven’t used “MustBeCalled” rather i used lamda specifically to assert the call that I have made, which is more justified for the cases where we exactly know the user code will behave. But, here goes a question that how we are going assert a mock call if we don’t know what item a user code may request for. In that case, we can combine the matchers with our assert calls like we do it for arrange: //setup - data  var order = new Order(TALISKER, 50);  var wareHouse = Mock.Create<IWarehouse>();    Mock.Arrange(() => wareHouse.HasInventory(TALISKER, Arg.Matches<int>( x => x <= 50))).Returns(true);    //exercise  order.Fill(wareHouse);    //verify state  Assert.True(order.IsFilled);    //verify interaction  Mock.Assert(() => wareHouse.HasInventory(Arg.Any<string>(), Arg.Matches<int>(x => x <= 50)));   Here, i have asserted a mock call for which i don’t know the item name,  but i know that number of items that user will request is less than 50.  This kind of expression based assertion is now possible with JustMock. We can extent this sample for properties as well, which will be covered shortly [in other posts]. In addition to just simple assertion, we can also use filters to limit to times a call has occurred or if ever occurred. Like for the first test code, we have one setup that is never invoked. For such, it is always valid to use the following assert call: Mock.Assert(() => wareHouse.Remove(Arg.Any<string>(), 100), Occurs.Never()); Or ,for warehouse.HasInventory we can do the following: Mock.Assert(() => wareHouse.HasInventory(Arg.Any<string>(), 0), Occurs.Once()); Or,  to be more specific, it’s even better with: Mock.Assert(() => wareHouse.HasInventory(Arg.Any<string>(), 0), Occurs.Exactly(1));   There are other filters  that you can apply here using AtMost, AtLeast and AtLeastOnce but I left those to the readers. You can try the above sample that is provided in the examples shipped with JustMock.Please, do check it out and feel free to ping me for any issues.   Enjoy!!

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  • LLBLGen Pro feature highlights: grouping model elements

    - by FransBouma
    (This post is part of a series of posts about features of the LLBLGen Pro system) When working with an entity model which has more than a few entities, it's often convenient to be able to group entities together if they belong to a semantic sub-model. For example, if your entity model has several entities which are about 'security', it would be practical to group them together under the 'security' moniker. This way, you could easily find them back, yet they can be left inside the complete entity model altogether so their relationships with entities outside the group are kept. In other situations your domain consists of semi-separate entity models which all target tables/views which are located in the same database. It then might be convenient to have a single project to manage the complete target database, yet have the entity models separate of each other and have them result in separate code bases. LLBLGen Pro can do both for you. This blog post will illustrate both situations. The feature is called group usage and is controllable through the project settings. This setting is supported on all supported O/R mapper frameworks. Situation one: grouping entities in a single model. This situation is common for entity models which are dense, so many relationships exist between all sub-models: you can't split them up easily into separate models (nor do you likely want to), however it's convenient to have them grouped together into groups inside the entity model at the project level. A typical example for this is the AdventureWorks example database for SQL Server. This database, which is a single catalog, has for each sub-group a schema, however most of these schemas are tightly connected with each other: adding all schemas together will give a model with entities which indirectly are related to all other entities. LLBLGen Pro's default setting for group usage is AsVisualGroupingMechanism which is what this situation is all about: we group the elements for visual purposes, it has no real meaning for the model nor the code generated. Let's reverse engineer AdventureWorks to an entity model. By default, LLBLGen Pro uses the target schema an element is in which is being reverse engineered, as the group it will be in. This is convenient if you already have categorized tables/views in schemas, like which is the case in AdventureWorks. Of course this can be switched off, or corrected on the fly. When reverse engineering, we'll walk through a wizard which will guide us with the selection of the elements which relational model data should be retrieved, which we can later on use to reverse engineer to an entity model. The first step after specifying which database server connect to is to select these elements. below we can see the AdventureWorks catalog as well as the different schemas it contains. We'll include all of them. After the wizard completes, we have all relational model data nicely in our catalog data, with schemas. So let's reverse engineer entities from the tables in these schemas. We select in the catalog explorer the schemas 'HumanResources', 'Person', 'Production', 'Purchasing' and 'Sales', then right-click one of them and from the context menu, we select Reverse engineer Tables to Entity Definitions.... This will bring up the dialog below. We check all checkboxes in one go by checking the checkbox at the top to mark them all to be added to the project. As you can see LLBLGen Pro has already filled in the group name based on the schema name, as this is the default and we didn't change the setting. If you want, you can select multiple rows at once and set the group name to something else using the controls on the dialog. We're fine with the group names chosen so we'll simply click Add to Project. This gives the following result:   (I collapsed the other groups to keep the picture small ;)). As you can see, the entities are now grouped. Just to see how dense this model is, I've expanded the relationships of Employee: As you can see, it has relationships with entities from three other groups than HumanResources. It's not doable to cut up this project into sub-models without duplicating the Employee entity in all those groups, so this model is better suited to be used as a single model resulting in a single code base, however it benefits greatly from having its entities grouped into separate groups at the project level, to make work done on the model easier. Now let's look at another situation, namely where we work with a single database while we want to have multiple models and for each model a separate code base. Situation two: grouping entities in separate models within the same project. To get rid of the entities to see the second situation in action, simply undo the reverse engineering action in the project. We still have the AdventureWorks relational model data in the catalog. To switch LLBLGen Pro to see each group in the project as a separate project, open the Project Settings, navigate to General and set Group usage to AsSeparateProjects. In the catalog explorer, select Person and Production, right-click them and select again Reverse engineer Tables to Entities.... Again check the checkbox at the top to mark all entities to be added and click Add to Project. We get two groups, as expected, however this time the groups are seen as separate projects. This means that the validation logic inside LLBLGen Pro will see it as an error if there's e.g. a relationship or an inheritance edge linking two groups together, as that would lead to a cyclic reference in the code bases. To see this variant of the grouping feature, seeing the groups as separate projects, in action, we'll generate code from the project with the two groups we just created: select from the main menu: Project -> Generate Source-code... (or press F7 ;)). In the dialog popping up, select the target .NET framework you want to use, the template preset, fill in a destination folder and click Start Generator (normal). This will start the code generator process. As expected the code generator has simply generated two code bases, one for Person and one for Production: The group name is used inside the namespace for the different elements. This allows you to add both code bases to a single solution and use them together in a different project without problems. Below is a snippet from the code file of a generated entity class. //... using System.Xml.Serialization; using AdventureWorks.Person; using AdventureWorks.Person.HelperClasses; using AdventureWorks.Person.FactoryClasses; using AdventureWorks.Person.RelationClasses; using SD.LLBLGen.Pro.ORMSupportClasses; namespace AdventureWorks.Person.EntityClasses { //... /// <summary>Entity class which represents the entity 'Address'.<br/><br/></summary> [Serializable] public partial class AddressEntity : CommonEntityBase //... The advantage of this is that you can have two code bases and work with them separately, yet have a single target database and maintain everything in a single location. If you decide to move to a single code base, you can do so with a change of one setting. It's also useful if you want to keep the groups as separate models (and code bases) yet want to add relationships to elements from another group using a copy of the entity: you can simply reverse engineer the target table to a new entity into a different group, effectively making a copy of the entity. As there's a single target database, changes made to that database are reflected in both models which makes maintenance easier than when you'd have a separate project for each group, with its own relational model data. Conclusion LLBLGen Pro offers a flexible way to work with entities in sub-models and control how the sub-models end up in the generated code.

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  • How can a code editor effectively hint at code nesting level - without using indentation?

    - by pgfearo
    I've written an XML text editor that provides 2 view options for the same XML text, one indented (virtually), the other left-justified. The motivation for the left-justified view is to help users 'see' the whitespace characters they're using for indentation of plain-text or XPath code without interference from indentation that is an automated side-effect of the XML context. I want to provide visual clues (in the non-editable part of the editor) for the left-justified mode that will help the user, but without getting too elaborate. I tried just using connecting lines, but that seemed too busy. The best I've come up with so far is shown in a mocked up screenshot of the editor below, but I'm seeking better/simpler alternatives (that don't require too much code). [Edit] Taking the heatmap idea (from: @jimp) I get this and 3 alternatives - labelled a, b and c: The following section describes the accepted answer as a proposal, bringing together ideas from a number of other answers and comments. As this question is now community wiki, please feel free to update this. NestView The name for this idea which provides a visual method to improve the readability of nested code without using indentation. Contour Lines The name for the differently shaded lines within the NestView The image above shows the NestView used to help visualise an XML snippet. Though XML is used for this illustration, any other code syntax that uses nesting could have been used for this illustration. An Overview: The contour lines are shaded (as in a heatmap) to convey nesting level The contour lines are angled to show when a nesting level is being either opened or closed. A contour line links the start of a nesting level to the corresponding end. The combined width of contour lines give a visual impression of nesting level, in addition to the heatmap. The width of the NestView may be manually resizable, but should not change as the code changes. Contour lines can either be compressed or truncated to keep acheive this. Blank lines are sometimes used code to break up text into more digestable chunks. Such lines could trigger special behaviour in the NestView. For example the heatmap could be reset or a background color contour line used, or both. One or more contour lines associated with the currently selected code can be highlighted. The contour line associated with the selected code level would be emphasized the most, but other contour lines could also 'light up' in addition to help highlight the containing nested group Different behaviors (such as code folding or code selection) can be associated with clicking/double-clicking on a Contour Line. Different parts of a contour line (leading, middle or trailing edge) may have different dynamic behaviors associated. Tooltips can be shown on a mouse hover event over a contour line The NestView is updated continously as the code is edited. Where nesting is not well-balanced assumptions can be made where the nesting level should end, but the associated temporary contour lines must be highlighted in some way as a warning. Drag and drop behaviors of Contour Lines can be supported. Behaviour may vary according to the part of the contour line being dragged. Features commonly found in the left margin such as line numbering and colour highlighting for errors and change state could overlay the NestView. Additional Functionality The proposal addresses a range of additional issues - many are outside the scope of the original question, but a useful side-effect. Visually linking the start and end of a nested region The contour lines connect the start and end of each nested level Highlighting the context of the currently selected line As code is selected, the associated nest-level in the NestView can be highlighted Differentiating between code regions at the same nesting level In the case of XML different hues could be used for different namespaces. Programming languages (such as c#) support named regions that could be used in a similar way. Dividing areas within a nesting area into different visual blocks Extra lines are often inserted into code to aid readability. Such empty lines could be used to reset the saturation level of the NestView's contour lines. Multi-Column Code View Code without indentation makes the use of a multi-column view more effective because word-wrap or horizontal scrolling is less likely to be required. In this view, once code has reach the bottom of one column, it flows into the next one: Usage beyond merely providing a visual aid As proposed in the overview, the NestView could provide a range of editing and selection features which would be broadly in line with what is expected from a TreeView control. The key difference is that a typical TreeView node has 2 parts: an expander and the node icon. A NestView contour line can have as many as 3 parts: an opener (sloping), a connector (vertical) and a close (sloping). On Indentation The NestView presented alongside non-indented code complements, but is unlikely to replace, the conventional indented code view. It's likely that any solutions adopting a NestView, will provide a method to switch seamlessly between indented and non-indented code views without affecting any of the code text itself - including whitespace characters. One technique for the indented view would be 'Virtual Formatting' - where a dynamic left-margin is used in lieu of tab or space characters. The same nesting-level data used to dynamically render the NestView could also used for the more conventional-looking indented view. Printing Indentation will be important for the readability of printed code. Here, the absence of tab/space characters and a dynamic left-margin means that the text can wrap at the right-margin and still maintain the integrity of the indented view. Line numbers can be used as visual markers that indicate where code is word-wrapped and also the exact position of indentation: Screen Real-Estate: Flat Vs Indented Addressing the question of whether the NestView uses up valuable screen real-estate: Contour lines work well with a width the same as the code editor's character width. A NestView width of 12 character widths can therefore accommodate 12 levels of nesting before contour lines are truncated/compressed. If an indented view uses 3 character-widths for each nesting level then space is saved until nesting reaches 4 levels of nesting, after this nesting level the flat view has a space-saving advantage that increases with each nesting level. Note: A minimum indentation of 4 character widths is often recommended for code, however XML often manages with less. Also, Virtual Formatting permits less indentation to be used because there's no risk of alignment issues A comparison of the 2 views is shown below: Based on the above, its probably fair to conclude that view style choice will be based on factors other than screen real-estate. The one exception is where screen space is at a premium, for example on a Netbook/Tablet or when multiple code windows are open. In these cases, the resizable NestView would seem to be a clear winner. Use Cases Examples of real-world examples where NestView may be a useful option: Where screen real-estate is at a premium a. On devices such as tablets, notepads and smartphones b. When showing code on websites c. When multiple code windows need to be visible on the desktop simultaneously Where consistent whitespace indentation of text within code is a priority For reviewing deeply nested code. For example where sub-languages (e.g. Linq in C# or XPath in XSLT) might cause high levels of nesting. Accessibility Resizing and color options must be provided to aid those with visual impairments, and also to suit environmental conditions and personal preferences: Compatability of edited code with other systems A solution incorporating a NestView option should ideally be capable of stripping leading tab and space characters (identified as only having a formatting role) from imported code. Then, once stripped, the code could be rendered neatly in both the left-justified and indented views without change. For many users relying on systems such as merging and diff tools that are not whitespace-aware this will be a major concern (if not a complete show-stopper). Other Works: Visualisation of Overlapping Markup Published research by Wendell Piez, dated from 2004, addresses the issue of the visualisation of overlapping markup, specifically LMNL. This includes SVG graphics with significant similarities to the NestView proposal, as such, they are acknowledged here. The visual differences are clear in the images (below), the key functional distinction is that NestView is intended only for well-nested XML or code, whereas Wendell Piez's graphics are designed to represent overlapped nesting. The graphics above were reproduced - with kind permission - from http://www.piez.org Sources: Towards Hermenutic Markup Half-steps toward LMNL

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  • API Message Localization

    - by Jesse Taber
    In my post, “Keep Localizable Strings Close To Your Users” I talked about the internationalization and localization difficulties that can arise when you sprinkle static localizable strings throughout the different logical layers of an application. The main point of that post is that you should have your localizable strings reside as close to the user-facing modules of your application as possible. For example, if you’re developing an ASP .NET web forms application all of the localizable strings should be kept in .resx files that are associated with the .aspx views of the application. In this post I want to talk about how this same concept can be applied when designing and developing APIs. An API Facilitates Machine-to-Machine Interaction You can typically think about a web, desktop, or mobile application as a collection “views” or “screens” through which users interact with the underlying logic and data. The application can be designed based on the assumption that there will be a human being on the other end of the screen working the controls. You are designing a machine-to-person interaction and the application should be built in a way that facilitates the user’s clear understanding of what is going on. Dates should be be formatted in a way that the user will be familiar with, messages should be presented in the user’s preferred language, etc. When building an API, however, there are no screens and you can’t make assumptions about who or what is on the other end of each call. An API is, by definition, a machine-to-machine interaction. A machine-to-machine interaction should be built in a way that facilitates a clear and unambiguous understanding of what is going on. Dates and numbers should be formatted in predictable and standard ways (e.g. ISO 8601 dates) and messages should be presented in machine-parseable formats. For example, consider an API for a time tracking system that exposes a resource for creating a new time entry. The JSON for creating a new time entry for a user might look like: 1: { 2: "userId": 4532, 3: "startDateUtc": "2012-10-22T14:01:54.98432Z", 4: "endDateUtc": "2012-10-22T11:34:45.29321Z" 5: }   Note how the parameters for start and end date are both expressed as ISO 8601 compliant dates in UTC. Using a date format like this in our API leaves little room for ambiguity. It’s also important to note that using ISO 8601 dates is a much, much saner thing than the \/Date(<milliseconds since epoch>)\/ nonsense that is sometimes used in JSON serialization. Probably the most important thing to note about the JSON snippet above is the fact that the end date comes before the start date! The API should recognize that and disallow the time entry from being created, returning an error to the caller. You might inclined to send a response that looks something like this: 1: { 2: "errors": [ {"message" : "The end date must come after the start date"}] 3: }   While this may seem like an appropriate thing to do there are a few problems with this approach: What if there is a user somewhere on the other end of the API call that doesn’t speak English?  What if the message provided here won’t fit properly within the UI of the application that made the API call? What if the verbiage of the message isn’t consistent with the rest of the application that made the API call? What if there is no user directly on the other end of the API call (e.g. this is a batch job uploading time entries once per night unattended)? The API knows nothing about the context from which the call was made. There are steps you could take to given the API some context (e.g.allow the caller to send along a language code indicating the language that the end user speaks), but that will only get you so far. As the designer of the API you could make some assumptions about how the API will be called, but if we start making assumptions we could very easily make the wrong assumptions. In this situation it’s best to make no assumptions and simply design the API in such a way that the caller has the responsibility to convey error messages in a manner that is appropriate for the context in which the error was raised. You would work around some of these problems by allowing callers to add metadata to each request describing the context from which the call is being made (e.g. accepting a ‘locale’ parameter denoting the desired language), but that will add needless clutter and complexity. It’s better to keep the API simple and push those context-specific concerns down to the caller whenever possible. For our very simple time entry example, this can be done by simply changing our error message response to look like this: 1: { 2: "errors": [ {"code": 100}] 3: }   By changing our error error from exposing a string to a numeric code that is easily parseable by another application, we’ve placed all of the responsibility for conveying the actual meaning of the error message on the caller. It’s best to have the caller be responsible for conveying this meaning because the caller understands the context much better than the API does. Now the caller can see error code 100, know that it means that the end date submitted falls before the start date and take appropriate action. Now all of the problems listed out above are non-issues because the caller can simply translate the error code of ‘100’ into the proper action and message for the current context. The numeric code representation of the error is a much better way to facilitate the machine-to-machine interaction that the API is meant to facilitate. An API Does Have Human Users While APIs should be built for machine-to-machine interaction, people still need to wire these interactions together. As a programmer building a client application that will consume the time entry API I would find it frustrating to have to go dig through the API documentation every time I encounter a new error code (assuming the documentation exists and is accurate). The numeric error code approach hurts the discoverability of the API and makes it painful to integrate with. We can help ease this pain by merging our two approaches: 1: { 2: "errors": [ {"code": 100, "message" : "The end date must come after the start date"}] 3: }   Now we have an easily parseable numeric error code for the machine-to-machine interaction that the API is meant to facilitate and a human-readable message for programmers working with the API. The human-readable message here is not intended to be viewed by end-users of the API and as such is not really a “localizable string” in my opinion. We could opt to expose a locale parameter for all API methods and store translations for all error messages, but that’s a lot of extra effort and overhead that doesn’t add a lot real value to the API. I might be a bit of an “ugly American”, but I think it’s probably fine to have the API return English messages when the target for those messages is a programmer. When resources are limited (which they always are), I’d argue that you’re better off hard-coding these messages in English and putting more effort into building more useful features, improving security, tweaking performance, etc.

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  • readonly keyword

    - by nmarun
    This is something new that I learned about the readonly keyword. Have a look at the following class: 1: public class MyClass 2: { 3: public string Name { get; set; } 4: public int Age { get; set; } 5:  6: private readonly double Delta; 7:  8: public MyClass() 9: { 10: Initializer(); 11: } 12:  13: public MyClass(string name = "", int age = 0) 14: { 15: Name = name; 16: Age = age; 17: Initializer(); 18: } 19:  20: private void Initializer() 21: { 22: Delta = 0.2; 23: } 24: } I have a couple of public properties and a private readonly member. There are two constructors – one that doesn’t take any parameters and the other takes two parameters to initialize the public properties. I’m also calling the Initializer method in both constructors to initialize the readonly member. Now when I build this, the code breaks and the Error window says: “A readonly field cannot be assigned to (except in a constructor or a variable initializer)” Two things after I read this message: It’s such a negative statement. I’d prefer something like: “A readonly field can be assigned to (or initialized) only in a constructor or through a variable initializer” But in my defense, I AM assigning it in a constructor (only indirectly). All I’m doing is creating a method that does it and calling it in a constructor. Turns out, .net was not ‘frameworked’ this way. We need to have the member initialized directly in the constructor. If you have multiple constructors, you can just use the ‘this’ keyword on all except the default constructors to call the default constructor. This default constructor can then initialize your readonly members. This will ensure you’re not repeating the code in multiple places. A snippet of what I’m talking can be seen below: 1: public class Person 2: { 3: public int UniqueNumber { get; set; } 4: public string Name { get; set; } 5: public int Age { get; set; } 6: public DateTime DateOfBirth { get; set; } 7: public string InvoiceNumber { get; set; } 8:  9: private readonly string Alpha; 10: private readonly int Beta; 11: private readonly double Delta; 12: private readonly double Gamma; 13:  14: public Person() 15: { 16: Alpha = "FDSA"; 17: Beta = 2; 18: Delta = 3.0; 19: Gamma = 0.0989; 20: } 21:  22: public Person(int uniqueNumber) : this() 23: { 24: UniqueNumber = uniqueNumber; 25: } 26: } See the syntax in line 22 and you’ll know what I’m talking about. So the default constructor gets called before the one in line 22. These are known as constructor initializers and they allow one constructor to call another. The other ‘myth’ I had about readonly members is that you can set it’s value only once. This was busted as well (I recall Adam and Jamie’s show). Say you’ve initialized the readonly member through a variable initializer. You can over-write this value in any of the constructors any number of times. 1: public class Person 2: { 3: public int UniqueNumber { get; set; } 4: public string Name { get; set; } 5: public int Age { get; set; } 6: public DateTime DateOfBirth { get; set; } 7: public string InvoiceNumber { get; set; } 8:  9: private readonly string Alpha = "asdf"; 10: private readonly int Beta = 15; 11: private readonly double Delta = 0.077; 12: private readonly double Gamma = 1.0; 13:  14: public Person() 15: { 16: Alpha = "FDSA"; 17: Beta = 2; 18: Delta = 3.0; 19: Gamma = 0.0989; 20: } 21:  22: public Person(int uniqueNumber) : this() 23: { 24: UniqueNumber = uniqueNumber; 25: Beta = 3; 26: } 27:  28: public Person(string name, DateTime dob) : this() 29: { 30: Name = name; 31: DateOfBirth = dob; 32:  33: Alpha = ";LKJ"; 34: Gamma = 0.0898; 35: } 36:  37: public Person(int uniqueNumber, string name, int age, DateTime dob, string invoiceNumber) : this() 38: { 39: UniqueNumber = uniqueNumber; 40: Name = name; 41: Age = age; 42: DateOfBirth = dob; 43: InvoiceNumber = invoiceNumber; 44:  45: Alpha = "QWER"; 46: Beta = 5; 47: Delta = 1.0; 48: Gamma = 0.0; 49: } 50: } In the above example, every constructor over-writes the values for the readonly members. This is perfectly valid. There is a possibility that based on the way the object is instantiated, the readonly member will have a different value. Well, that’s all I have for today and read this as it’s on a related topic.

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  • ASP.NET MVC 2 throws exception for ‘favicon.ico’

    - by nmarun
    I must be on fire or something – third blog in 2 days… awesome! Before I begin, in case you’re wondering, favicon.ico is the small image that appears to the left of your web address, once the page loads. In order to learn more about MVC or any thing for that matter, it’s better to look at the source itself. Since MVC is open source (at least some part of it is), I started looking at the source code that’s available for download. While doing so, I hit Steve Sanderson’s blog site where he explains in great detail the way to debug your app using ASP.NET MVC source code. For those who are not aware, Steve Sanderson’s book - Pro ASP.NET MVC Framework, is one of the best books to learn about MVC. Alrighty, I followed the article and I hit F5 to debug the default / unchanged MVC project. I put a breakpoint in the DefaultControllerFactory.cs, CreateController() method. To know a little more about this class and the method, read this. Sure enough, the control stopped at the breakpoint and I hit F5 again and the page rendered just fine. But then what’s this? The breakpoint was hit again, as if something else was being requested. I now hovered my mouse over the ‘controllerName’ parameter and it says – favicon.ico. This by itself was more than enough for me to raise my eye-brows, but what happened next just took the ground below my feet. Oh, oh, I’m sorry I’m just typing, no code, no image, so here are a couple of screen captures. The first one shows the request for the Home controller; I get ‘Home’ when I hover over the parameter: And here’s the one that shows the same for call for ‘favicon.ico’. So, I step through the code and when the control reaches line 91 – GetControllerInstance() method, I step in. This is when I had the ‘ground-losing’ experience. Wow, an exception is being thrown for this file and that too in RTM. For some reason MVC thinks, this as a controller and tries to run it through the MvcHandler and it hits this snag. So it seems like this will happen for any MVC 2 site and this did not happen for me in the previous version of MVC. Before I get to how to resolve it, here’s another way of reproducing this exception. Revert back all your changes that you did as mentioned in Steve’s blog above. Now, add a class to your MVC project and call it say, MyControllerFactory and let this inherit from DefaultControllerFactory class. (Read this for details on the DefaultControllerFactory class is and how it is used in a different context). Add an override for the CreateController() method and for the sake of this blog, just copy the same content from the DefaultControllerFactory class. The last step is to tell your MVC app to use the MyControllerFactory class instead of the default one. To do this, go to your Global.asax.cs file and add line 6 of the snippet below: 1: protected void Application_Start() 2: { 3: AreaRegistration.RegisterAllAreas(); 4:   5: RegisterRoutes(RouteTable.Routes); 6: ControllerBuilder.Current.SetControllerFactory(new MyControllerFactory()); 7: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now, you’re ready to reproduce the issue. Just F5 the project and when you hit the overridden CreateController() method for the second time, this is what it looks like for me: And continuing further gives me the same exception. I believe this is something that MS should fix, as not having ‘favicon.ico’ file will be common for most of the applications. So I think the when you create an MVC project, line 6 should be added by default by Visual Studio itself: 1: public class MvcApplication : System.Web.HttpApplication 2: { 3: public static void RegisterRoutes(RouteCollection routes) 4: { 5: routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); 6: routes.IgnoreRoute("favicon.ico"); 7:   8: routes.MapRoute( 9: "Default", // Route name 10: "{controller}/{action}/{id}", // URL with parameters 11: new { controller = "Home", action = "Index", id = UrlParameter.Optional } // Parameter defaults 12: ); 13: } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } There it is, that’s the solution to avoid the exception altogether. I tried this both IE8 and Firefox browsers and was able to successfully reproduce the error. Hope someone will look at this issue and find a fix. Just before I finish up, I found another ‘bug’, if you want to call it, with Visual Studio 2008. Remember how you could change what browser you want your application to run in by just right clicking on the .aspx file and choosing ‘Browse with…’? Seems like that’s missing when you’re working with an MVC project. In order to test the above bug in the other browser, I had to load a classic ASP.NET project, change the settings and then run my MVC project. Felt kinda ‘icky’, for lack of a better word.

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  • Working With Extended Events

    - by Fatherjack
    SQL Server 2012 has made working with Extended Events (XE) pretty simple when it comes to what sessions you have on your servers and what options you have selected and so forth but if you are like me then you still have some SQL Server instances that are 2008 or 2008 R2. For those servers there is no built-in way to view the Extended Event sessions in SSMS. I keep coming up against the same situations – Where are the xel log files? What events, actions or predicates are set for the events on the server? What sessions are there on the server already? I got tired of this being a perpetual question and wrote some TSQL to save as a snippet in SQL Prompt so that these details are permanently only a couple of clicks away. First, some history. If you just came here for the code skip down a few paragraphs and it’s all there. If you want a little time to reminisce about SQL Server then stick with me through the next paragraph or two. We are in a bit of a cross-over period currently, there are many versions of SQL Server but I would guess that SQL Server 2008, 2008 R2 and 2012 comprise the majority of installations. With each of these comes a set of management tools, of which SQL Server Management Studio (SSMS) is one. In 2008 and 2008 R2 Extended Events made their first appearance and there was no way to work with them in the SSMS interface. At some point the Extended Events guru Jonathan Kehayias (http://www.sqlskills.com/blogs/jonathan/) created the SQL Server 2008 Extended Events SSMS Addin which is really an excellent tool to ease XE session administration. This addin will install in SSMS 2008 or 2008R2 but not SSMS 2012. If you use a compatible version of SSMS then I wholly recommend downloading and using it to make your work with XE much easier. If you have SSMS 2012 installed, and there is no reason not to as it will let you work with all versions of SQL Server, then you cannot install this addin. If you are working with SQL Server 2012 then SSMS 2012 has built in functionality to manage XE sessions – this functionality does not apply for 2008 or 2008 R2 instances though. This means you are somewhat restricted and have to use TSQL to manage XE sessions on older versions of SQL Server. OK, those of you that skipped ahead for the code, you need to start from here: So, you are working with SSMS 2012 but have a SQL Server that is an earlier version that needs an XE session created or you think there is a session created but you aren’t sure, or you know it’s there but can’t remember if it is running and where the output is going. How do you find out? Well, none of the information is hidden as such but it is a bit of a wrangle to locate it and it isn’t a lot of code that is unlikely to remain in your memory. I have created two pieces of code. The first examines the SYS.Server_Event_… management views in combination with the SYS.DM_XE_… management views to give the name of all sessions that exist on the server, regardless of whether they are running or not and two pieces of TSQL code. One piece will alter the state of the session: if the session is running then the code will stop the session if executed and vice versa. The other piece of code will drop the selected session. If the session is running then the code will stop it first. Do not execute the DROP code unless you are sure you have the Create code to hand. It will be dropped from the server without a second chance to change your mind. /**************************************************************/ /***   To locate and describe event sessions on a server    ***/ /***                                                        ***/ /***   Generates TSQL to start/stop/drop sessions           ***/ /***                                                        ***/ /***        Jonathan Allen - @fatherjack                    ***/ /***                 June 2013                                ***/ /***                                                        ***/ /**************************************************************/ SELECT  [EES].[name] AS [Session Name - all sessions] ,         CASE WHEN [MXS].[name] IS NULL THEN ISNULL([MXS].[name], 'Stopped')              ELSE 'Running'         END AS SessionState ,         CASE WHEN [MXS].[name] IS NULL              THEN ISNULL([MXS].[name],                          'ALTER EVENT SESSION [' + [EES].[name]                          + '] ON SERVER STATE = START;')              ELSE 'ALTER EVENT SESSION [' + [EES].[name]                   + '] ON SERVER STATE = STOP;'         END AS ALTER_SessionState ,         CASE WHEN [MXS].[name] IS NULL              THEN ISNULL([MXS].[name],                          'DROP EVENT SESSION [' + [EES].[name]                          + '] ON SERVER; -- This WILL drop the session. It will no longer exist. Don't do it unless you are certain you can recreate it if you need it.')              ELSE 'ALTER EVENT SESSION [' + [EES].[name]                   + '] ON SERVER STATE = STOP; ' + CHAR(10)                   + '-- DROP EVENT SESSION [' + [EES].[name]                   + '] ON SERVER; -- This WILL stop and drop the session. It will no longer exist. Don't do it unless you are certain you can recreate it if you need it.'         END AS DROP_Session FROM    [sys].[server_event_sessions] AS EES         LEFT JOIN [sys].[dm_xe_sessions] AS MXS ON [EES].[name] = [MXS].[name] WHERE   [EES].[name] NOT IN ( 'system_health', 'AlwaysOn_health' ) ORDER BY SessionState GO I have excluded the system_health and AlwaysOn sessions as I don’t want to accidentally execute the drop script for these sessions that are created as part of the SQL Server installation. It is possible to recreate the sessions but that is a whole lot of aggravation I’d rather avoid. The second piece of code gathers details of running XE sessions only and provides information on the Events being collected, any predicates that are set on those events, the actions that are set to be collected, where the collected information is being logged and if that logging is to a file target, where that file is located. /**********************************************/ /***    Running Session summary                ***/ /***                                        ***/ /***    Details key values of XE sessions     ***/ /***    that are in a running state            ***/ /***                                        ***/ /***        Jonathan Allen - @fatherjack    ***/ /***        June 2013                        ***/ /***                                        ***/ /**********************************************/ SELECT  [EES].[name] AS [Session Name - running sessions] ,         [EESE].[name] AS [Event Name] ,         COALESCE([EESE].[predicate], 'unfiltered') AS [Event Predicate Filter(s)] ,         [EESA].[Action] AS [Event Action(s)] ,         [EEST].[Target] AS [Session Target(s)] ,         ISNULL([EESF].[value], 'No file target in use') AS [File_Target_UNC] -- select * FROM    [sys].[server_event_sessions] AS EES         INNER JOIN [sys].[dm_xe_sessions] AS MXS ON [EES].[name] = [MXS].[name]         INNER JOIN [sys].[server_event_session_events] AS [EESE] ON [EES].[event_session_id] = [EESE].[event_session_id]         LEFT JOIN [sys].[server_event_session_fields] AS EESF ON ( [EES].[event_session_id] = [EESF].[event_session_id]                                                               AND [EESF].[name] = 'filename'                                                               )         CROSS APPLY ( SELECT    STUFF(( SELECT  ', ' + sest.name                                         FROM    [sys].[server_event_session_targets]                                                 AS SEST                                         WHERE   [EES].[event_session_id] = [SEST].[event_session_id]                                       FOR                                         XML PATH('')                                       ), 1, 2, '') AS [Target]                     ) AS EEST         CROSS APPLY ( SELECT    STUFF(( SELECT  ', ' + [sesa].NAME                                         FROM    [sys].[server_event_session_actions]                                                 AS sesa                                         WHERE   [sesa].[event_session_id] = [EES].[event_session_id]                                       FOR                                         XML PATH('')                                       ), 1, 2, '') AS [Action]                     ) AS EESA WHERE   [EES].[name] NOT IN ( 'system_health', 'AlwaysOn_health' ) /*Optional to exclude 'out-of-the-box' traces*/ I hope that these scripts are useful to you and I would be obliged if you would keep my name in the script comments. I have no problem with you using it in production or personal circumstances, however it has no warranty or guarantee. Don’t use it unless you understand it and are happy with what it is going to do. I am not ever responsible for the consequences of executing this script on your servers.

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  • Prevent your Silverlight XAP file from caching in your browser.

    - by mbcrump
    If you work with Silverlight daily then you have run into this problem. Your XAP file has been cached in your browser and you have to empty your browser cache to resolve it. If your using Google Chrome then you typically do the following: Go to Options –> Clear Browsing History –> Empty the Cache and finally click Clear Browsing data. As you can see, this is a lot of unnecessary steps. It is even worse when you have a customer that says, “I can’t see the new features you just implemented!” and you realize it’s a cached xap problem.  I have been struggling with a way to prevent my XAP file from caching inside of a browser for a while now and decided to implement the following solution. If the Visual Studio Debugger is attached then add a unique query string to the source param to force the XAP file to be refreshed. If the Visual Studio Debugger is not attached then add the source param as Visual Studio generates it. This is also in case I forget to remove the above code in my production environment. I want the ASP.NET code to be inline with my .ASPX page. (I do not want a separate code behind .cs page or .vb page attached to the .aspx page.) Below is an example of the hosting code generated when you create a new Silverlight project. As a quick refresher, the hard coded param name = “source” specifies the location of your XAP file.  <form id="form1" runat="server" style="height:100%"> <div id="silverlightControlHost"> <object data="data:application/x-silverlight-2," type="application/x-silverlight-2" width="100%" height="100%"> <param name="source" value="ClientBin/SilverlightApplication2.xap"/> <param name="onError" value="onSilverlightError" /> <param name="background" value="white" /> <param name="minRuntimeVersion" value="4.0.50826.0" /> <param name="autoUpgrade" value="true" /> <a href="http://go.microsoft.com/fwlink/?LinkID=149156&v=4.0.50826.0" style="text-decoration:none"> <img src="http://go.microsoft.com/fwlink/?LinkId=161376" alt="Get Microsoft Silverlight" style="border-style:none"/> </a> </object><iframe id="_sl_historyFrame" style="visibility:hidden;height:0px;width:0px;border:0px"></iframe></div> </form> We are going to use a little bit of inline ASP.NET to generate the param name = source dynamically to prevent the XAP file from caching. Lets look at the completed solution: <form id="form1" runat="server" style="height:100%"> <div id="silverlightControlHost"> <object data="data:application/x-silverlight-2," type="application/x-silverlight-2" width="100%" height="100%"> <% string strSourceFile = @"ClientBin/SilverlightApplication2.xap"; string param; if (System.Diagnostics.Debugger.IsAttached) //Debugger Attached - Refresh the XAP file. param = "<param name=\"source\" value=\"" + strSourceFile + "?" + DateTime.Now.Ticks + "\" />"; else { //Production Mode param = "<param name=\"source\" value=\"" + strSourceFile + "\" />"; } Response.Write(param); %> <param name="onError" value="onSilverlightError" /> <param name="background" value="white" /> <param name="minRuntimeVersion" value="4.0.50826.0" /> <param name="autoUpgrade" value="true" /> <a href="http://go.microsoft.com/fwlink/?LinkID=149156&v=4.0.50826.0" style="text-decoration:none"> <img src="http://go.microsoft.com/fwlink/?LinkId=161376" alt="Get Microsoft Silverlight" style="border-style:none"/> </a> </object><iframe id="_sl_historyFrame" style="visibility:hidden;height:0px;width:0px;border:0px"></iframe></div> </form> We add the location to our XAP file to strSourceFile and if the debugger is attached then it will append DateTime.Now.Ticks to the XAP file source and force the browser to download the .XAP. If you view the page source of your Silverlight Application then you can verify it worked properly by looking at the param name = “source” tag as shown below. <param name="source" value="ClientBin/SilverlightApplication2.xap?634299001187160148" /> If the debugger is not attached then it will use the standard source tag as shown below. <param name="source" value="ClientBin/SilverlightApplication2.xap"/> At this point you may be asking, How do I prevent my XAP file from being cached on my production app? Well, you have two easy options: 1) I really don’t recommend this approach but you can force the XAP to be refreshed everytime with the following code snippet.  <param name="source" value="ClientBin/SilverlightApplication2.xap?<%=Guid.NewGuid().ToString() %>"/> NOTE: You could also substitute the “Guid.NewGuid().ToString() for anything that create a random field. (I used DateTime.Now.Ticks earlier). 2) Another solution that I like even better involves checking the XAP Creation Date and appending it to the param name = source. This method was described by Lars Holm Jenson. <% string strSourceFile = @"ClientBin/SilverlightApplication2.xap"; string param; if (System.Diagnostics.Debugger.IsAttached) param = "<param name=\"source\" value=\"" + strSourceFile + "\" />"; else { string xappath = HttpContext.Current.Server.MapPath(@"") + @"\" + strSourceFile; DateTime xapCreationDate = System.IO.File.GetLastWriteTime(xappath); param = "<param name=\"source\" value=\"" + strSourceFile + "?ignore=" + xapCreationDate.ToString() + "\" />"; } Response.Write(param); %> As you can see, this problem has been solved. It will work with all web browsers and stubborn proxy servers that are caching your .XAP. If you enjoyed this article then check out my blog for others like this. You may also want to subscribe to my blog or follow me on Twitter.   Subscribe to my feed

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  • Using Open MQ as an Oracle CEP Event Source

    - by seth.white
    I helped an Oracle CEP customer recently who wanted to use Open MQ has an event source for their Oracle CEP application.  In this case, the Oracle CEP application was being used to provide monitoring for an electronic commerce website, however, the steps for configuring Open MQ are entirely independent of the application logic. I thought I would list the configuration steps in a blog post in case they might help others in the future. Note that although the Oracle CEP documentation states that only WebLogic and Tibco JMS are "officially" supported, any JMS implementation that provides a Java client should work with Oracle CEP. The first step is to add an adapter to the application's EPN. This can be done in the usual way, using the Eclipse IDE. The end result is something like the following bit of configuration in the application's Spring application context. Note that the provider attribute value of 'jms-inbound' specifies that the out-of-the-box JMS adapter is being used. <wlevs:adapter id="helloworldAdapter" provider="jms-inbound"> </wlevs:adapter>   Next, configure the inbound adapter so that it can connect to Open MQ in the Oracle CEP configuration file (config.xml). The snippet below provides an example of what this configuration should look like. The exact values specified for jndi-provider-url, jndi-factory, connection-jndi-name, destination-jndi-name elements will depend on your Open MQ configuration.  For example , if the name of your Open MQ topic destination is 'ElectronicCommerceTopic', then you would specify that as the destination-jndi-name.  The name of your Open MQ connection factory goes in the connection-jndi-name element. In my simple example, I also specify in event-type element so that the out-of-the-box JMS adapter will attempt to automatically convert incoming messages to events of type HelloWorldEvent. In a more complex application, one would configure a custom converter on the JMS adapter to convert from messages to events.  The Oracle CEP 11.1.3 documentation describes how to do this.   <jms-adapter> <name>helloworldAdapter</name> <event-type>HelloWorldEvent</event-type> <jndi-provider-url>file:///C:/Temp</jndi-provider-url> <jndi-factory>com.sun.jndi.fscontext.RefFSContextFactory</jndi-factory> <connection-jndi-name>YourJMSConnectionFactoryName</connection-jndi-name> <destination-jndi-name>YourJMSDestinationName</destination-jndi-name> </jms-adapter>   Finally, one needs to package the client-side Open MQ jars so that the classes that they contain are available to the Oracle CEP runtime. The recommended way for doing this in the Oracle CEP 11.1.3 release is to package the classes as a library module or simply place them in the application bundle.  The advantage of deploying the classes as a library module is that they are available to any application that wants to connect to Open MQ. In my case, I packaged the classes in my application bundle. A best practice when you want to include additional jars in your application bundle is to create a 'lib' directory in your Eclipse project and then copy the required jars into that directory.  Then, use the support that Eclipse provides to add the jars to the bundle classpath (which makes the classes part of your application in the same way that regular application classes are), and export all of the classes from your application bundle so that they are available to the Oracle CEP server runtime.  The screenshot below Illustrates how this is done in Eclipse.  The bundle classpath contains two Open MQ jars and all packages in the jars are exported.     Finally, import the javax.jms and javax.naming packages into the application module as these are needed by the Open MQ classes. The screenshot below shows the complete list of package imports for my sample application.       Once you have completed these steps, you should be able to build and deploy your application and begin receiving inbound messages from Open MQ. 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  • tile_static, tile_barrier, and tiled matrix multiplication with C++ AMP

    - by Daniel Moth
    We ended the previous post with a mechanical transformation of the C++ AMP matrix multiplication example to the tiled model and in the process introduced tiled_index and tiled_grid. This is part 2. tile_static memory You all know that in regular CPU code, static variables have the same value regardless of which thread accesses the static variable. This is in contrast with non-static local variables, where each thread has its own copy. Back to C++ AMP, the same rules apply and each thread has its own value for local variables in your lambda, whereas all threads see the same global memory, which is the data they have access to via the array and array_view. In addition, on an accelerator like the GPU, there is a programmable cache, a third kind of memory type if you'd like to think of it that way (some call it shared memory, others call it scratchpad memory). Variables stored in that memory share the same value for every thread in the same tile. So, when you use the tiled model, you can have variables where each thread in the same tile sees the same value for that variable, that threads from other tiles do not. The new storage class for local variables introduced for this purpose is called tile_static. You can only use tile_static in restrict(direct3d) functions, and only when explicitly using the tiled model. What this looks like in code should be no surprise, but here is a snippet to confirm your mental image, using a good old regular C array // each tile of threads has its own copy of locA, // shared among the threads of the tile tile_static float locA[16][16]; Note that tile_static variables are scoped and have the lifetime of the tile, and they cannot have constructors or destructors. tile_barrier In amp.h one of the types introduced is tile_barrier. You cannot construct this object yourself (although if you had one, you could use a copy constructor to create another one). So how do you get one of these? You get it, from a tiled_index object. Beyond the 4 properties returning index objects, tiled_index has another property, barrier, that returns a tile_barrier object. The tile_barrier class exposes a single member, the method wait. 15: // Given a tiled_index object named t_idx 16: t_idx.barrier.wait(); 17: // more code …in the code above, all threads in the tile will reach line 16 before a single one progresses to line 17. Note that all threads must be able to reach the barrier, i.e. if you had branchy code in such a way which meant that there is a chance that not all threads could reach line 16, then the code above would be illegal. Tiled Matrix Multiplication Example – part 2 So now that we added to our understanding the concepts of tile_static and tile_barrier, let me obfuscate rewrite the matrix multiplication code so that it takes advantage of tiling. Before you start reading this, I suggest you get a cup of your favorite non-alcoholic beverage to enjoy while you try to fully understand the code. 01: void MatrixMultiplyTiled(vector<float>& vC, const vector<float>& vA, const vector<float>& vB, int M, int N, int W) 02: { 03: static const int TS = 16; 04: array_view<const float,2> a(M, W, vA); 05: array_view<const float,2> b(W, N, vB); 06: array_view<writeonly<float>,2> c(M,N,vC); 07: parallel_for_each(c.grid.tile< TS, TS >(), 08: [=] (tiled_index< TS, TS> t_idx) restrict(direct3d) 09: { 10: int row = t_idx.local[0]; int col = t_idx.local[1]; 11: float sum = 0.0f; 12: for (int i = 0; i < W; i += TS) { 13: tile_static float locA[TS][TS], locB[TS][TS]; 14: locA[row][col] = a(t_idx.global[0], col + i); 15: locB[row][col] = b(row + i, t_idx.global[1]); 16: t_idx.barrier.wait(); 17: for (int k = 0; k < TS; k++) 18: sum += locA[row][k] * locB[k][col]; 19: t_idx.barrier.wait(); 20: } 21: c[t_idx.global] = sum; 22: }); 23: } Notice that all the code up to line 9 is the same as per the changes we made in part 1 of tiling introduction. If you squint, the body of the lambda itself preserves the original algorithm on lines 10, 11, and 17, 18, and 21. The difference being that those lines use new indexing and the tile_static arrays; the tile_static arrays are declared and initialized on the brand new lines 13-15. On those lines we copy from the global memory represented by the array_view objects (a and b), to the tile_static vanilla arrays (locA and locB) – we are copying enough to fit a tile. Because in the code that follows on line 18 we expect the data for this tile to be in the tile_static storage, we need to synchronize the threads within each tile with a barrier, which we do on line 16 (to avoid accessing uninitialized memory on line 18). We also need to synchronize the threads within a tile on line 19, again to avoid the race between lines 14, 15 (retrieving the next set of data for each tile and overwriting the previous set) and line 18 (not being done processing the previous set of data). Luckily, as part of the awesome C++ AMP debugger in Visual Studio there is an option that helps you find such races, but that is a story for another blog post another time. May I suggest reading the next section, and then coming back to re-read and walk through this code with pen and paper to really grok what is going on, if you haven't already? Cool. Why would I introduce this tiling complexity into my code? Funny you should ask that, I was just about to tell you. There is only one reason we tiled our extent, had to deal with finding a good tile size, ensure the number of threads we schedule are correctly divisible with the tile size, had to use a tiled_index instead of a normal index, and had to understand tile_barrier and to figure out where we need to use it, and double the size of our lambda in terms of lines of code: the reason is to be able to use tile_static memory. Why do we want to use tile_static memory? Because accessing tile_static memory is around 10 times faster than accessing the global memory on an accelerator like the GPU, e.g. in the code above, if you can get 150GB/second accessing data from the array_view a, you can get 1500GB/second accessing the tile_static array locA. And since by definition you are dealing with really large data sets, the savings really pay off. We have seen tiled implementations being twice as fast as their non-tiled counterparts. Now, some algorithms will not have performance benefits from tiling (and in fact may deteriorate), e.g. algorithms that require you to go only once to global memory will not benefit from tiling, since with tiling you already have to fetch the data once from global memory! Other algorithms may benefit, but you may decide that you are happy with your code being 150 times faster than the serial-version you had, and you do not need to invest to make it 250 times faster. Also algorithms with more than 3 dimensions, which C++ AMP supports in the non-tiled model, cannot be tiled. Also note that in future releases, we may invest in making the non-tiled model, which already uses tiling under the covers, go the extra step and use tile_static memory on your behalf, but it is obviously way to early to commit to anything like that, and we certainly don't do any of that today. Comments about this post by Daniel Moth welcome at the original blog.

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  • Tracing Silex from PHP to the OS with DTrace

    - by cj
    In this blog post I show the full stack tracing of Brendan Gregg's php_syscolors.d script in the DTrace Toolkit. The Toolkit contains a dozen very useful PHP DTrace scripts and many more scripts for other languages and the OS. For this example, I'll trace the PHP micro framework Silex, which was the topic of the second of two talks by Dustin Whittle at a recent SF PHP Meetup. His slides are at Silex: From Micro to Full Stack. Installing DTrace and PHP The php_syscolors.d script uses some static PHP probes and some kernel probes. For Oracle Linux I discussed installing DTrace and PHP in DTrace PHP Using Oracle Linux 'playground' Pre-Built Packages. On other platforms with DTrace support, follow your standard procedures to enable DTrace and load the correct providers. The sdt and systrace providers are required in addition to fasttrap. On Oracle Linux, I loaded the DTrace modules like: # modprobe fasttrap # modprobe sdt # modprobe systrace # chmod 666 /dev/dtrace/helper Installing the DTrace Toolkit I download DTraceToolkit-0.99.tar.gz and extracted it: $ tar -zxf DTraceToolkit-0.99.tar.gz The PHP scripts are in the Php directory and examples in the Examples directory. Installing Silex I downloaded the "fat" Silex .tgz file from the download page and extracted it: $ tar -zxf silex_fat.tgz I changed the demonstration silex/web/index.php so I could use the PHP development web server: <?php // web/index.php $filename = __DIR__.preg_replace('#(\?.*)$#', '', $_SERVER['REQUEST_URI']); if (php_sapi_name() === 'cli-server' && is_file($filename)) { return false; } require_once __DIR__.'/../vendor/autoload.php'; $app = new Silex\Application(); //$app['debug'] = true; $app->get('/hello', function() { return 'Hello!'; }); $app->run(); ?> Running DTrace The php_syscolors.d script uses the -Z option to dtrace, so it can be started before PHP, i.e. when there are zero of the requested probes available to be traced. I ran DTrace like: # cd DTraceToolkit-0.99/Php # ./php_syscolors.d Next, I started the PHP developer web server in a second terminal: $ cd silex $ php -S localhost:8080 -t web web/index.php At this point, the web server is idle, waiting for requests. DTrace is idle, waiting for the probes in php_syscolors.d to be fired, at which time the action associated with each probe will run. I then loaded the demonstration page in a browser: http://localhost:8080/hello When the request was fulfilled and the simple output of "Hello" was displayed, I ^C'd php and dtrace in their terminals to stop them. DTrace output over a thousand lines long had been generated. Here is one snippet from when run() was invoked: C PID/TID DELTA(us) FILE:LINE TYPE -- NAME ... 1 4765/4765 21 Application.php:487 func -> run 1 4765/4765 29 ClassLoader.php:182 func -> loadClass 1 4765/4765 17 ClassLoader.php:198 func -> findFile 1 4765/4765 31 ":- syscall -> access 1 4765/4765 26 ":- syscall <- access 1 4765/4765 16 ClassLoader.php:198 func <- findFile 1 4765/4765 25 ":- syscall -> newlstat 1 4765/4765 15 ":- syscall <- newlstat 1 4765/4765 13 ":- syscall -> newlstat 1 4765/4765 13 ":- syscall <- newlstat 1 4765/4765 22 ":- syscall -> newlstat 1 4765/4765 14 ":- syscall <- newlstat 1 4765/4765 15 ":- syscall -> newlstat 1 4765/4765 60 ":- syscall <- newlstat 1 4765/4765 13 ":- syscall -> newlstat 1 4765/4765 13 ":- syscall <- newlstat 1 4765/4765 20 ":- syscall -> open 1 4765/4765 16 ":- syscall <- open 1 4765/4765 26 ":- syscall -> newfstat 1 4765/4765 12 ":- syscall <- newfstat 1 4765/4765 17 ":- syscall -> newfstat 1 4765/4765 12 ":- syscall <- newfstat 1 4765/4765 12 ":- syscall -> newfstat 1 4765/4765 12 ":- syscall <- newfstat 1 4765/4765 20 ":- syscall -> mmap 1 4765/4765 14 ":- syscall <- mmap 1 4765/4765 3201 ":- syscall -> mmap 1 4765/4765 27 ":- syscall <- mmap 1 4765/4765 1233 ":- syscall -> munmap 1 4765/4765 53 ":- syscall <- munmap 1 4765/4765 15 ":- syscall -> close 1 4765/4765 13 ":- syscall <- close 1 4765/4765 34 Request.php:32 func -> main 1 4765/4765 22 Request.php:32 func <- main 1 4765/4765 31 ClassLoader.php:182 func <- loadClass 1 4765/4765 33 Request.php:249 func -> createFromGlobals 1 4765/4765 29 Request.php:198 func -> __construct 1 4765/4765 24 Request.php:218 func -> initialize 1 4765/4765 26 ClassLoader.php:182 func -> loadClass 1 4765/4765 89 ClassLoader.php:198 func -> findFile 1 4765/4765 43 ":- syscall -> access ... The output shows PHP functions being called and returning (and where they are located) and which system calls the PHP functions in turn invoked. The time each line took from the previous one is displayed in the third column. The first column is the CPU number. In this example, the process was always on CPU 1 so the output is naturally ordered without requiring post-processing, or the D script requiring to be modified to display a time stamp. On a terminal, the output of php_syscolors.d is color-coded according to whether each function is a PHP or system one, hence the file name. Summary With one tool, I was able to trace the interaction of a user application with the operating system. I was able to do this to an application running "live" in a web context. The DTrace Toolkit provides a very handy repository of DTrace information. Even though the PHP scripts were created in the time frame of the original PHP DTrace PECL extension, which only had PHP function entry and return probes, the scripts provide core examples for custom investigation and resolution scripts. You can easily adapt the ideas and and create scripts using the other PHP static probes, which are listed in the PHP Manual. Because DTrace is "always on", you can take advantage of it to resolve development questions or fix production situations.

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  • Instructions on how to configure a WebLogic Cluster and use it with Oracle Http Server

    - by Laurent Goldsztejn
    On October 17th I delivered a webcast on WebLogic Clustering that included a demo with Apache as the proxy server.  I realized that many steps are needed to set up the configuration I used during the demo.  The purpose of this article is to go through these steps to show how quickly and easily one can define a new cluster and then proxy requests via an Oracle Http Server (OHS). The domain configuration wizard offers the option to create a cluster.  The administration console or WLST, the Weblogic scripting tool can also be used to define a new cluster.  It can be created at any time but the servers that will participate in it cannot be in a running state. Cluster Creation using the configuration wizard Network and architecture requirements need to be considered while choosing between unicast and multicast. Multicast Vs. Unicast with WebLogic Clustering is of great help to make the best decision between the two messaging modes.  In addition, Configure Cluster offers details on each single field displayed above. After this initial configuration page, individual servers could be assigned to this newly created cluster although servers can be added later to the cluster.  What is not recommended is for the Admin server to participate in a cluster as the main purpose of the Admin server is to perform the bulk of the processing for the domain.  Servers need to stop before being assigned to a cluster.  There is also no minimum number of servers that have to participate in the cluster. At this point the configuration should be done and the cluster created successfully.  This can easily be verified from the console. Each clustered managed server can be launched to join the cluster.   At startup the following messages should be logged for each clustered managed server: <Notice> <WeblogicServer> <BEA-000365> <Server state changed to STARTING> <Notice> <Cluster> <BEA-000197> <Listening for announcements from cluster using messaging_mode cluster messaging> <Notice> <Cluster> <BEA-000133> <Waiting to synchronize with other running members of cluster_name>  It's time to try sending requests to the cluster and we will do this with the help of Oracle Http Server to play the role of a proxy server to demonstrate load balancing.  Proxy Server configuration  The first step is to download Weblogic Server Web Server Plugin that will enhance the web server by handling requests aimed at being sent to the Weblogic cluster.  For our test Oracle Http Server (OHS) will be used.  However plug-ins are also available for Apache Http server, Microsoft Internet Information Server (IIS), Oracle iPlanet Webserver or even WebLogic Server with the HttpClusterServlet. Once OHS is installed on the system, the configuration file, mod_wl_ohs.conf, will need to be altered to include Weblogic proxy specifics. First of all, add the following directive to instruct Apache to load the Weblogic shared object module extracted from the plugins file just downloaded. LoadModule weblogic_module modules/mod_wl_ohs.so and then create an IfModule directive to encapsulate the following location block so that proxy will be enabled by path (each request including /wls will be directed directly to the WebLogic Cluster).  You could also proxy requests by MIME type using MatchExpression in the Location block. <IfModule weblogic_module> <Location /wls>    SetHandler weblogic-handler    PathTrim /wls    WebLogicCluster MS1_URL:port,MS2_URL:port    Debug ON    WLLogFile        c:/tmp/global_proxy.log     WLTempDir        "c:/myTemp"    DebugConfigInfo  On </Location> </IfModule> SetHandler specifies the handler for the plug-in module  PathTrim will instruct the plug-in to trim /w ls from the URL before forwarding the request to the cluster. The list of WebLogic Servers defined in WeblogicCluster could contain a mixed set of clustered and single servers.  However, the dynamic list returned for this parameter will only contain valid clustered servers and may contain more servers if not all clustered servers are listed in WeblogicCluster. Testing proxy and load balancing It's time to start OHS web server which should at this point be configured correctly to proxy requests to the clustered servers.  By default round-robin is the load balancing strategy set by WebLogic. Testing the load balancing can be easily done by disabling cookies on your browser given that a request containing a cookie attempts to connect to the primary server. If that attempt fails, the plug-in attempts to make a connection to the next available server in the list in a round-robin fashion.  With cookies enabled, you could use two different browsers to test the load balancing with a JSP page that contains the following: <%@ page contentType="text/html; charset=iso-8859-1" language="java"  %>  <%  String path = request.getContextPath();   String getProtocol=request.getScheme();   String getDomain=request.getServerName();   String getPort=Integer.toString(request.getLocalPort());   String getPath = getProtocol+"://"+getDomain+":"+getPort+path+"/"; %> <html> <body> Receiving Server <%=getPath%> </body> </html>  Assuming that you name the JSP page Test.jsp and the webapp that contains it TestApp, your browsers should open the following URL: http://localhost/wls/TestApp/Test.jsp  Each browser should connect to a different clustered server and this simple JSP should confirm that.  The webapp that contains the JSP needs to be deployed to the cluster. You can also verify that the load is correctly balanced by looking at the proxy log file.  Each request generates a set of log entries that starts with : timestamp ================New Request: Each request is associated with a primary server and a secondary server if one is available.  For our test request, the following entries should appear in the log as well:Using Uri /wls/TestApp/Test.jsp After trimming path: '/TestApp/Test.jsp' The final request string is '/TestApp/Test.jsp' If an exception occurs, it should also be logged in the proxy log file with the prefix:timestamp *******Exception type   WeblogicBridgeConfig DebugConfigInfo enables runtime statistics and the production of configuration information.  For security purposes, this parameter should be turned off in production. http://webserver_host:port/path/xyz.jsp?__WebLogicBridgeConfig will display a proxy bridge page detailing the plugin configuration followed by runtime statistics which could help in diagnosing issues along with the analyzing of the proxy log file.  In our example the url would be: http://localhost/wls/TestApp/Test.jsp?__WebLogicBridgeConfig  Here is how the top section of the screen can look like: The bottom part of the page contains runtime statistics, here is a snippet of it (unrelated with the previous JSP example).   This entire plugin configuration should be very similar with other web servers, what varies is the name of the proxy server configuration file. So, as you can see, it only takes a few minutes to configure a Weblogic cluster and get servers to join it. 

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  • REST to Objects in C#

    RESTful interfaces for web services are all the rage for many Web 2.0 sites.  If you want to consume these in a very simple fashion, LINQ to XML can do the job pretty easily in C#.  If you go searching for help on this, youll find a lot of incomplete solutions and fairly large toolkits and frameworks (guess how I know this) this quick article is meant to be a no fluff just stuff approach to making this work. POCO Objects Lets assume you have a Model that you want to suck data into from a RESTful web service.  Ideally this is a Plain Old CLR Object, meaning it isnt infected with any persistence or serialization goop.  It might look something like this: public class Entry { public int Id; public int UserId; public DateTime Date; public float Hours; public string Notes; public bool Billable;   public override string ToString() { return String.Format("[{0}] User: {1} Date: {2} Hours: {3} Notes: {4} Billable {5}", Id, UserId, Date, Hours, Notes, Billable); } } Not that this isnt a completely trivial object.  Lets look at the API for the service.  RESTful HTTP Service In this case, its TickSpots API, with the following sample output: <?xml version="1.0" encoding="UTF-8"?> <entries type="array"> <entry> <id type="integer">24</id> <task_id type="integer">14</task_id> <user_id type="integer">3</user_id> <date type="date">2008-03-08</date> <hours type="float">1.00</hours> <notes>Had trouble with tribbles.</notes> <billable>true</billable> # Billable is an attribute inherited from the task <billed>true</billed> # Billed is an attribute to track whether the entry has been invoiced <created_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</created_at> <updated_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</updated_at> # The following attributes are derived and provided for informational purposes: <user_email>[email protected]</user_email> <task_name>Remove converter assembly</task_name> <sum_hours type="float">2.00</sum_hours> <budget type="float">10.00</budget> <project_name>Realign dilithium crystals</project_name> <client_name>Starfleet Command</client_name> </entry> </entries> Im assuming in this case that I dont necessarily care about all of the data fields the service is returning I just need some of them for my applications purposes.  Thus, you can see there are more elements in the <entry> XML than I have in my Entry class. Get The XML with C# The next step is to get the XML.  The following snippet does the heavy lifting once you pass it the appropriate URL: protected XElement GetResponse(string uri) { var request = WebRequest.Create(uri) as HttpWebRequest; request.UserAgent = ".NET Sample"; request.KeepAlive = false;   request.Timeout = 15 * 1000;   var response = request.GetResponse() as HttpWebResponse;   if (request.HaveResponse == true && response != null) { var reader = new StreamReader(response.GetResponseStream()); return XElement.Parse(reader.ReadToEnd()); } throw new Exception("Error fetching data."); } This is adapted from the Yahoo Developer article on Web Service REST calls.  Once you have the XML, the last step is to get the data back as your POCO. Use LINQ-To-XML to Deserialize POCOs from XML This is done via the following code: public IEnumerable<Entry> List(DateTime startDate, DateTime endDate) { string additionalParameters = String.Format("start_date={0}&end_date={1}", startDate.ToShortDateString(), endDate.ToShortDateString()); string uri = BuildUrl("entries", additionalParameters);   XElement elements = GetResponse(uri);   var entries = from e in elements.Elements() where e.Name.LocalName == "entry" select new Entry { Id = int.Parse(e.Element("id").Value), UserId = int.Parse(e.Element("user_id").Value), Date = DateTime.Parse(e.Element("date").Value), Hours = float.Parse(e.Element("hours").Value), Notes = e.Element("notes").Value, Billable = bool.Parse(e.Element("billable").Value) }; return entries; }   For completeness, heres the BuildUrl method for my TickSpot API wrapper: // Change these to your settings protected const string projectDomain = "DOMAIN.tickspot.com"; private const string authParams = "[email protected]&password=MyTickSpotPassword";   protected string BuildUrl(string apiMethod, string additionalParams) { if (projectDomain.Contains("DOMAIN")) { throw new ApplicationException("You must update your domain in ProjectRepository.cs."); } if (authParams.Contains("MyTickSpotPassword")) { throw new ApplicationException("You must update your email and password in ProjectRepository.cs."); } return string.Format("https://{0}/api/{1}?{2}&{3}", projectDomain, apiMethod, authParams, additionalParams); } Thats it!  Now go forth and consume XML and map it to classes you actually want to work with.  Have fun! Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • MapRedux - PowerShell and Big Data

    - by Dittenhafer Solutions
    MapRedux – #PowerShell and #Big Data Have you been hearing about “big data”, “map reduce” and other large scale computing terms over the past couple of years and been curious to dig into more detail? Have you read some of the Apache Hadoop online documentation and unfortunately concluded that it wasn't feasible to setup a “test” hadoop environment on your machine? More recently, I have read about some of Microsoft’s work to enable Hadoop on the Azure cloud. Being a "Microsoft"-leaning technologist, I am more inclinded to be successful with experimentation when on the Windows platform. Of course, it is not that I am "religious" about one set of technologies other another, but rather more experienced. Anyway, within the past couple of weeks I have been thinking about PowerShell a bit more as the 2012 PowerShell Scripting Games approach and it occured to me that PowerShell's support for Windows Remote Management (WinRM), and some other inherent features of PowerShell might lend themselves particularly well to a simple implementation of the MapReduce framework. I fired up my PowerShell ISE and started writing just to see where it would take me. Quite simply, the ScriptBlock feature combined with the ability of Invoke-Command to create remote jobs on networked servers provides much of the plumbing of a distributed computing environment. There are some limiting factors of course. Microsoft provided some default settings which prevent PowerShell from taking over a network without administrative approval first. But even with just one adjustment, a given Windows-based machine can become a node in a MapReduce-style distributed computing environment. Ok, so enough introduction. Let's talk about the code. First, any machine that will participate as a remote "node" will need WinRM enabled for remote access, as shown below. This is not exactly practical for hundreds of intended nodes, but for one (or five) machines in a test environment it does just fine. C:> winrm quickconfig WinRM is not set up to receive requests on this machine. The following changes must be made: Set the WinRM service type to auto start. Start the WinRM service. Make these changes [y/n]? y Alternatively, you could take the approach described in the Remotely enable PSRemoting post from the TechNet forum and use PowerShell to create remote scheduled tasks that will call Enable-PSRemoting on each intended node. Invoke-MapRedux Moving on, now that you have one or more remote "nodes" enabled, you can consider the actual Map and Reduce algorithms. Consider the following snippet: $MyMrResults = Invoke-MapRedux -MapReduceItem $Mr -ComputerName $MyNodes -DataSet $dataset -Verbose Invoke-MapRedux takes an instance of a MapReduceItem which references the Map and Reduce scriptblocks, an array of computer names which are the remote nodes, and the initial data set to be processed. As simple as that, you can start working with concepts of big data and the MapReduce paradigm. Now, how did we get there? I have published the initial version of my PsMapRedux PowerShell Module on GitHub. The PsMapRedux module provides the Invoke-MapRedux function described above. Feel free to browse the underlying code and even contribute to the project! In a later post, I plan to show some of the inner workings of the module, but for now let's move on to how the Map and Reduce functions are defined. Map Both the Map and Reduce functions need to follow a prescribed prototype. The prototype for a Map function in the MapRedux module is as follows. A simple scriptblock that takes one PsObject parameter and returns a hashtable. It is important to note that the PsObject $dataset parameter is a MapRedux custom object that has a "Data" property which offers an array of data to be processed by the Map function. $aMap = { Param ( [PsObject] $dataset ) # Indicate the job is running on the remote node. Write-Host ($env:computername + "::Map"); # The hashtable to return $list = @{}; # ... Perform the mapping work and prepare the $list hashtable result with your custom PSObject... # ... The $dataset has a single 'Data' property which contains an array of data rows # which is a subset of the originally submitted data set. # Return the hashtable (Key, PSObject) Write-Output $list; } Reduce Likewise, with the Reduce function a simple prototype must be followed which takes a $key and a result $dataset from the MapRedux's partitioning function (which joins the Map results by key). Again, the $dataset is a MapRedux custom object that has a "Data" property as described in the Map section. $aReduce = { Param ( [object] $key, [PSObject] $dataset ) Write-Host ($env:computername + "::Reduce - Count: " + $dataset.Data.Count) # The hashtable to return $redux = @{}; # Return Write-Output $redux; } All Together Now When everything is put together in a short example script, you implement your Map and Reduce functions, query for some starting data, build the MapReduxItem via New-MapReduxItem and call Invoke-MapRedux to get the process started: # Import the MapRedux and SQL Server providers Import-Module "MapRedux" Import-Module “sqlps” -DisableNameChecking # Query the database for a dataset Set-Location SQLSERVER:\sql\dbserver1\default\databases\myDb $query = "SELECT MyKey, Date, Value1 FROM BigData ORDER BY MyKey"; Write-Host "Query: $query" $dataset = Invoke-SqlCmd -query $query # Build the Map function $MyMap = { Param ( [PsObject] $dataset ) Write-Host ($env:computername + "::Map"); $list = @{}; foreach($row in $dataset.Data) { # Write-Host ("Key: " + $row.MyKey.ToString()); if($list.ContainsKey($row.MyKey) -eq $true) { $s = $list.Item($row.MyKey); $s.Sum += $row.Value1; $s.Count++; } else { $s = New-Object PSObject; $s | Add-Member -Type NoteProperty -Name MyKey -Value $row.MyKey; $s | Add-Member -type NoteProperty -Name Sum -Value $row.Value1; $list.Add($row.MyKey, $s); } } Write-Output $list; } $MyReduce = { Param ( [object] $key, [PSObject] $dataset ) Write-Host ($env:computername + "::Reduce - Count: " + $dataset.Data.Count) $redux = @{}; $count = 0; foreach($s in $dataset.Data) { $sum += $s.Sum; $count += 1; } # Reduce $redux.Add($s.MyKey, $sum / $count); # Return Write-Output $redux; } # Create the item data $Mr = New-MapReduxItem "My Test MapReduce Job" $MyMap $MyReduce # Array of processing nodes... $MyNodes = ("node1", "node2", "node3", "node4", "localhost") # Run the Map Reduce routine... $MyMrResults = Invoke-MapRedux -MapReduceItem $Mr -ComputerName $MyNodes -DataSet $dataset -Verbose # Show the results Set-Location C:\ $MyMrResults | Out-GridView Conclusion I hope you have seen through this article that PowerShell has a significant infrastructure available for distributed computing. While it does take some code to expose a MapReduce-style framework, much of the work is already done and PowerShell could prove to be the the easiest platform to develop and run big data jobs in your corporate data center, potentially in the Azure cloud, or certainly as an academic excerise at home or school. Follow me on Twitter to stay up to date on the continuing progress of my Powershell MapRedux module, and thanks for reading! Daniel

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  • Loaded OBJ Model Will Not Display in OpenGL / C++ Project

    - by Drake Summers
    I have been experimenting with new effects in game development. The programs I have written have been using generic shapes for the visuals. I wanted to test the effects on something a bit more complex, and wrote a resource loader for Wavefront OBJ files. I started with a simple cube in blender, exported it to an OBJ file with just vertices and triangulated faces, and used it to test the resource loader. I could not get the mesh to show up in my application. The loader never gave me any errors, so I wrote a snippet to loop through my vertex and index arrays that were returned from the loader. The data is exactly the way it is supposed to be. So I simplified the OBJ file by editing it directly to just show a front facing square. Still, nothing is displayed in the application. And don't worry, I did check to make sure that I decreased the value of each index by one while importing the OBJ. - BEGIN EDIT I also tested using glDrawArrays(GL_TRIANGLES, 0, 3 ); to draw the first triangle and it worked! So the issue could be in the binding of the VBO/IBO items. END EDIT - INDEX/VERTEX ARRAY OUTPUT: GLOBALS AND INITIALIZATION FUNCTION: GLuint program; GLint attrib_coord3d; std::vector<GLfloat> vertices; std::vector<GLushort> indices; GLuint vertexbuffer, indexbuffer; GLint uniform_mvp; int initialize() { if (loadModel("test.obj", vertices, indices)) { GLfloat myverts[vertices.size()]; copy(vertices.begin(), vertices.end(), myverts); GLushort myinds[indices.size()]; copy(indices.begin(), indices.end(), myinds); glGenBuffers(1, &vertexbuffer); glBindBuffer(GL_ARRAY_BUFFER, vertexbuffer); glBufferData(GL_ARRAY_BUFFER, sizeof(myverts), myverts, GL_STATIC_DRAW); glGenBuffers(1, &indexbuffer); glBindBuffer(GL_ARRAY_BUFFER, indexbuffer); glBufferData(GL_ELEMENT_ARRAY_BUFFER, sizeof(myinds), myinds, GL_STATIC_DRAW); // OUTPUT DATA FROM NEW ARRAYS TO CONSOLE // ERROR HANDLING OMITTED FOR BREVITY } GLint link_result = GL_FALSE; GLuint vert_shader, frag_shader; if ((vert_shader = create_shader("tri.v.glsl", GL_VERTEX_SHADER)) == 0) return 0; if ((frag_shader = create_shader("tri.f.glsl", GL_FRAGMENT_SHADER)) == 0) return 0; program = glCreateProgram(); glAttachShader(program, vert_shader); glAttachShader(program, frag_shader); glLinkProgram(program); glGetProgramiv(program, GL_LINK_STATUS, &link_result); // ERROR HANDLING OMITTED FOR BREVITY const char* attrib_name; attrib_name = "coord3d"; attrib_coord3d = glGetAttribLocation(program, attrib_name); // ERROR HANDLING OMITTED FOR BREVITY const char* uniform_name; uniform_name = "mvp"; uniform_mvp = glGetUniformLocation(program, uniform_name); // ERROR HANDLING OMITTED FOR BREVITY return 1; } RENDERING FUNCTION: glm::mat4 model = glm::translate(glm::mat4(1.0f), glm::vec3(0.0, 0.0, -4.0)); glm::mat4 view = glm::lookAt(glm::vec3(0.0, 0.0, 4.0), glm::vec3(0.0, 0.0, 3.0), glm::vec3(0.0, 1.0, 0.0)); glm::mat4 projection = glm::perspective(45.0f, 1.0f*(screen_width/screen_height), 0.1f, 10.0f); glm::mat4 mvp = projection * view * model; int size; glUseProgram(program); glUniformMatrix4fv(uniform_mvp, 1, GL_FALSE, glm::value_ptr(mvp)); glClearColor(0.5, 0.5, 0.5, 1.0); glClear(GL_COLOR_BUFFER_BIT|GL_DEPTH_BUFFER_BIT); glEnableVertexAttribArray(attrib_coord3d); glBindBuffer(GL_ARRAY_BUFFER, vertexbuffer); glVertexAttribPointer(attrib_coord3d, 3, GL_FLOAT, GL_FALSE, 0, 0); glBindBuffer(GL_ELEMENT_ARRAY_BUFFER, indexbuffer); glGetBufferParameteriv(GL_ELEMENT_ARRAY_BUFFER, GL_BUFFER_SIZE, &size); glDrawElements(GL_TRIANGLES, size/sizeof(GLushort), GL_UNSIGNED_SHORT, 0); glDisableVertexAttribArray(attrib_coord3d); VERTEX SHADER: attribute vec3 coord3d; uniform mat4 mvp; void main(void) { gl_Position = mvp * vec4(coord3d, 1.0); } FRAGMENT SHADER: void main(void) { gl_FragColor[0] = 0.0; gl_FragColor[1] = 0.0; gl_FragColor[2] = 1.0; gl_FragColor[3] = 1.0; } OBJ RESOURCE LOADER: bool loadModel(const char * path, std::vector<GLfloat> &out_vertices, std::vector<GLushort> &out_indices) { std::vector<GLfloat> temp_vertices; std::vector<GLushort> vertexIndices; FILE * file = fopen(path, "r"); // ERROR HANDLING OMITTED FOR BREVITY while(1) { char lineHeader[128]; int res = fscanf(file, "%s", lineHeader); if (res == EOF) { break; } if (strcmp(lineHeader, "v") == 0) { float _x, _y, _z; fscanf(file, "%f %f %f\n", &_x, &_y, &_z ); out_vertices.push_back(_x); out_vertices.push_back(_y); out_vertices.push_back(_z); } else if (strcmp(lineHeader, "f") == 0) { unsigned int vertexIndex[3]; int matches = fscanf(file, "%d %d %d\n", &vertexIndex[0], &vertexIndex[1], &vertexIndex[2]); out_indices.push_back(vertexIndex[0] - 1); out_indices.push_back(vertexIndex[1] - 1); out_indices.push_back(vertexIndex[2] - 1); } else { ... } } // ERROR HANDLING OMITTED FOR BREVITY return true; } I can edit the question to provide any further info you may need. I attempted to provide everything of relevance and omit what may have been unnecessary. I'm hoping this isn't some really poor mistake, because I have been at this for a few days now. If anyone has any suggestions or advice on the matter, I look forward to hearing it. As a final note: I added some arrays into the code with manually entered data, and was able to display meshes by using those arrays instead of the generated ones. I do not understand!

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  • REST to Objects in C#

    RESTful interfaces for web services are all the rage for many Web 2.0 sites.  If you want to consume these in a very simple fashion, LINQ to XML can do the job pretty easily in C#.  If you go searching for help on this, youll find a lot of incomplete solutions and fairly large toolkits and frameworks (guess how I know this) this quick article is meant to be a no fluff just stuff approach to making this work. POCO Objects Lets assume you have a Model that you want to suck data into from a RESTful web service.  Ideally this is a Plain Old CLR Object, meaning it isnt infected with any persistence or serialization goop.  It might look something like this: public class Entry { public int Id; public int UserId; public DateTime Date; public float Hours; public string Notes; public bool Billable;   public override string ToString() { return String.Format("[{0}] User: {1} Date: {2} Hours: {3} Notes: {4} Billable {5}", Id, UserId, Date, Hours, Notes, Billable); } } Not that this isnt a completely trivial object.  Lets look at the API for the service.  RESTful HTTP Service In this case, its TickSpots API, with the following sample output: <?xml version="1.0" encoding="UTF-8"?> <entries type="array"> <entry> <id type="integer">24</id> <task_id type="integer">14</task_id> <user_id type="integer">3</user_id> <date type="date">2008-03-08</date> <hours type="float">1.00</hours> <notes>Had trouble with tribbles.</notes> <billable>true</billable> # Billable is an attribute inherited from the task <billed>true</billed> # Billed is an attribute to track whether the entry has been invoiced <created_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</created_at> <updated_at type="datetime">Tue, 07 Oct 2008 14:46:16 -0400</updated_at> # The following attributes are derived and provided for informational purposes: <user_email>[email protected]</user_email> <task_name>Remove converter assembly</task_name> <sum_hours type="float">2.00</sum_hours> <budget type="float">10.00</budget> <project_name>Realign dilithium crystals</project_name> <client_name>Starfleet Command</client_name> </entry> </entries> Im assuming in this case that I dont necessarily care about all of the data fields the service is returning I just need some of them for my applications purposes.  Thus, you can see there are more elements in the <entry> XML than I have in my Entry class. Get The XML with C# The next step is to get the XML.  The following snippet does the heavy lifting once you pass it the appropriate URL: protected XElement GetResponse(string uri) { var request = WebRequest.Create(uri) as HttpWebRequest; request.UserAgent = ".NET Sample"; request.KeepAlive = false;   request.Timeout = 15 * 1000;   var response = request.GetResponse() as HttpWebResponse;   if (request.HaveResponse == true && response != null) { var reader = new StreamReader(response.GetResponseStream()); return XElement.Parse(reader.ReadToEnd()); } throw new Exception("Error fetching data."); } This is adapted from the Yahoo Developer article on Web Service REST calls.  Once you have the XML, the last step is to get the data back as your POCO. Use LINQ-To-XML to Deserialize POCOs from XML This is done via the following code: public IEnumerable<Entry> List(DateTime startDate, DateTime endDate) { string additionalParameters = String.Format("start_date={0}&end_date={1}", startDate.ToShortDateString(), endDate.ToShortDateString()); string uri = BuildUrl("entries", additionalParameters);   XElement elements = GetResponse(uri);   var entries = from e in elements.Elements() where e.Name.LocalName == "entry" select new Entry { Id = int.Parse(e.Element("id").Value), UserId = int.Parse(e.Element("user_id").Value), Date = DateTime.Parse(e.Element("date").Value), Hours = float.Parse(e.Element("hours").Value), Notes = e.Element("notes").Value, Billable = bool.Parse(e.Element("billable").Value) }; return entries; }   For completeness, heres the BuildUrl method for my TickSpot API wrapper: // Change these to your settings protected const string projectDomain = "DOMAIN.tickspot.com"; private const string authParams = "[email protected]&password=MyTickSpotPassword";   protected string BuildUrl(string apiMethod, string additionalParams) { if (projectDomain.Contains("DOMAIN")) { throw new ApplicationException("You must update your domain in ProjectRepository.cs."); } if (authParams.Contains("MyTickSpotPassword")) { throw new ApplicationException("You must update your email and password in ProjectRepository.cs."); } return string.Format("https://{0}/api/{1}?{2}&{3}", projectDomain, apiMethod, authParams, additionalParams); } Thats it!  Now go forth and consume XML and map it to classes you actually want to work with.  Have fun! Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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