Search Results

Search found 17816 results on 713 pages for 'variable names'.

Page 649/713 | < Previous Page | 645 646 647 648 649 650 651 652 653 654 655 656  | Next Page >

  • Functional Adaptation

    - by Charles Courchaine
    In real life and OO programming we’re often faced with using adapters, DVI to VGA, 1/4” to 1/8” audio connections, 110V to 220V, wrapping an incompatible interface with a new one, and so on.  Where the adapter pattern is generally considered for interfaces and classes a similar technique can be applied to method signatures.  To be fair, this adaptation is generally used to reduce the number of parameters but I’m sure there are other clever possibilities to be had.  As Jan questioned in the last post, how can we use a common method to execute an action if the action has a differing number of parameters, going back to the greeting example it was suggested having an AddName method that takes a first and last name as parameters.  This is exactly what we’ll address in this post. Let’s set the stage with some review and some code changes.  First, our method that handles the setup/tear-down infrastructure for our WCF service: 1: private static TResult ExecuteGreetingFunc<TResult>(Func<IGreeting, TResult> theGreetingFunc) 2: { 3: IGreeting aGreetingService = null; 4: try 5: { 6: aGreetingService = GetGreetingChannel(); 7: return theGreetingFunc(aGreetingService); 8: } 9: finally 10: { 11: CloseWCFChannel((IChannel)aGreetingService); 12: } 13: } Our original AddName method: 1: private static string AddName(string theName) 2: { 3: return ExecuteGreetingFunc<string>(theGreetingService => theGreetingService.AddName(theName)); 4: } Our new AddName method: 1: private static int AddName(string firstName, string lastName) 2: { 3: return ExecuteGreetingFunc<int>(theGreetingService => theGreetingService.AddName(firstName, lastName)); 4: } Let’s change the AddName method, just a little bit more for this example and have it take the greeting service as a parameter. 1: private static int AddName(IGreeting greetingService, string firstName, string lastName) 2: { 3: return greetingService.AddName(firstName, lastName); 4: } The new signature of AddName using the Func delegate is now Func<IGreeting, string, string, int>, which can’t be used with ExecuteGreetingFunc as is because it expects Func<IGreeting, TResult>.  Somehow we have to eliminate the two string parameters before we can use this with our existing method.  This is where we need to adapt AddName to match what ExecuteGreetingFunc expects, and we’ll do so in the following progression. 1: Func<IGreeting, string, string, int> -> Func<IGreeting, string, int> 2: Func<IGreeting, string, int> -> Func<IGreeting, int>   For the first step, we’ll create a method using the lambda syntax that will “eliminate” the last name parameter: 1: string lastNameToAdd = "Smith"; 2: //Func<IGreeting, string, string, int> -> Func<IGreeting, string, int> 3: Func<IGreeting, string, int> addName = (greetingService, firstName) => AddName(greetingService, firstName, lastNameToAdd); The new addName method gets us one step close to the signature we need.  Let’s say we’re going to call this in a loop to add several names, we’ll take the final step from Func<IGreeting, string, int> -> Func<IGreeting, int> in line as a lambda passed to ExecuteGreetingFunc like so: 1: List<string> firstNames = new List<string>() { "Bob", "John" }; 2: int aID; 3: foreach (string firstName in firstNames) 4: { 5: //Func<IGreeting, string, int> -> Func<IGreeting, int> 6: aID = ExecuteGreetingFunc<int>(greetingService => addName(greetingService, firstName)); 7: Console.WriteLine(GetGreeting(aID)); 8: } If for some reason you needed to break out the lambda on line 6 you could replace it with 1: aID = ExecuteGreetingFunc<int>(ApplyAddName(addName, firstName)); and use this method: 1: private static Func<IGreeting, int> ApplyAddName(Func<IGreeting, string, int> addName, string lastName) 2: { 3: return greetingService => addName(greetingService, lastName); 4: } Splitting out a lambda into its own method is useful both in this style of coding as well as LINQ queries to improve the debugging experience.  It is not strictly necessary to break apart the steps & functions as was shown above; the lambda in line 6 (of the foreach example) could include both the last name and first name instead of being composed of two functions.  The process demonstrated above is one of partially applying functions, this could have also been done with Currying (also see Dustin Campbell’s excellent post on Currying for the canonical curried add example).  Matthew Podwysocki also has some good posts explaining both Currying and partial application and a follow up post that further clarifies the difference between Currying and partial application.  In either technique the ultimate goal is to reduce the number of parameters passed to a function.  Currying makes it a single parameter passed at each step, where partial application allows one to use multiple parameters at a time as we’ve done here.  This technique isn’t for everyone or every problem, but can be extremely handy when you need to adapt a call to something you don’t control.

    Read the article

  • WebLogic JDBC Use of Oracle Wallet for SSL

    - by Steve Felts
    Introduction Secure Sockets Layer (SSL) can be used to secure the connection between the middle tier “client”, WebLogic Server (WLS) in this case, and the Oracle database server.  Data between WLS and database can be encrypted.  The server can be authenticated so you have proof that the database can be trusted by validating a certificate from the server.  The client can be authenticated so that the database only accepts connections from clients that it trusts. Similar to the discussion in an earlier article about using the Oracle wallet for database credentials, the Oracle wallet can also be used with SSL to store the keys and certificates.  By using it correctly, clear text passwords can be eliminated from the JDBC configuration and client/server configuration can be simplified by sharing the wallet across multiple datasources. There is a very good Oracle Technical White Paper on using SSL with the Oracle thin driver at http://www.oracle.com/technetwork/database/enterprise-edition/wp-oracle-jdbc-thin-ssl-130128.pdf [LINK1].  The link http://www.oracle.com/technetwork/middleware/weblogic/index-087556.html [LINK2] describes how to use WebLogic Server with Oracle JDBC Driver SSL. The information in this article is a guide on what steps need to be taken in the variety of available options; use the links above for details. SSL from the driver to the database server is basically turned on by specifying a protocol of “tcps” in the URL.  However, there is a fair amount of setup needed.  Also remember that there is an overhead in performance. Creating the wallets The common use cases are 1. “data encryption and server-only authentication”, requiring just a trust store, or 2. “data encryption and authentication of both tiers” (client and server), requiring a trust store and a key store. It is recommended to use the auto-login wallet type so that clear text passwords are not needed in the datasource configuration to open the wallet.  The store type for an auto-login wallet is “SSO” (Single Sign On), not “JKS” or “PKCS12” as in [LINK2].  The file name is “cwallet.sso”. Wallets are created using the orapki tool.  They need to be created based on the usage (encryption and/or authentication).  This is discussed in detail in [LINK1] in Appendix B or in the Advanced Security Administrator’s Guide of the Database documentation. Database Server Configuration It is necessary to update the sqlnet.ora and listener.ora files with the directory location of the wallet using WALLET_LOCATION.  These files also indicate whether or not SSL_CLIENT_AUTHENTICATION is being used (true or false). The Oracle Listener must also be configured to use the TCPS protocol.  The recommended port is 2484. LISTENER = (ADDRESS_LIST= (ADDRESS=(PROTOCOL=tcps)(HOST=servername)(PORT=2484))) WebLogic Server Classpath The WebLogic Server CLASSPATH must have three additional security files. The files that need to be added to the WLS CLASSPATH are $MW_HOME/modules/com.oracle.osdt_cert_1.0.0.0.jar $MW_HOME/modules/com.oracle.osdt_core_1.0.0.0.jar $MW_HOME/modules/com.oracle.oraclepki_1.0.0.0.jar One way to do this is to add them to PRE_CLASSPATH environment variable for use with the standard WebLogic scripts. Setting the Oracle Security Provider It’s necessary to enable the Oracle PKI provider on the client side.  This can either be done statically by updating the java.security file under the JRE or dynamically by setting it in a WLS startup class using java.security.Security.insertProviderAt(new oracle.security.pki.OraclePKIProvider (), 3); See the full example of the startup class in [LINK2]. Datasource Configuration When creating a WLS datasource, set the PROTOCOL in the URL to tcps as in the following. jdbc:oracle:thin:@(DESCRIPTION=(ADDRESS=(PROTOCOL=tcps)(HOST=host)(PORT=port))(CONNECT_DATA=(SERVICE_NAME=myservice))) For encryption and server authentication, use the datasource connection properties: - javax.net.ssl.trustStore=location of wallet file on the client - javax.net.ssl.trustStoreType=”SSO” For client authentication, use the datasource connection properties: - javax.net.ssl.keyStore=location of wallet file on the client - javax.net.ssl.keyStoreType=”SSO” Note that the driver connection properties for the wallet require a file name, not a directory name. Active GridLink ONS over SSL For completeness, there is another SSL usage for WLS datasources.  The communication with the Oracle Notification Service (ONS) for load balancing information and node up/down events can use SSL also. Create an auto-login wallet and use the wallet on the client and server.  The following is a sample sequence to create a test wallet for use with ONS. orapki wallet create -wallet ons -auto_login -pwd ONS_Wallet orapki wallet add -wallet ons -dn "CN=ons_test,C=US" -keysize 1024 -self_signed -validity 9999 -pwd ONS_Wallet orapki wallet export -wallet ons -dn "CN=ons_test,C=US" -cert ons/cert.txt -pwd ONS_Wallet On the database server side, it’s necessary to define the walletfile directory in the file $CRS_HOME/opmn/conf/ons.config and run onsctl stop/start. When configuring an Active GridLink datasource, the connection to the ONS must be defined.  In addition to the host and port, the wallet file directory must be specified.  By not giving a password, a SSO wallet is assumed. Summary To use SSL with the Oracle thin driver without any clear text passwords, use an SSO Oracle Wallet.  SSL support in the Oracle thin driver is available starting in 10g Release 2.

    Read the article

  • WIF, ADFS 2 and WCF&ndash;Part 5: Service Client (more Flexibility with WSTrustChannelFactory)

    - by Your DisplayName here!
    See the previous posts first. WIF includes an API to manually request tokens from a token service. This gives you more control over the request and more flexibility since you can use your own token caching scheme instead of being bound to the channel object lifetime. The API is straightforward. You first request a token from the STS and then use that token to create a channel to the relying party service. I’d recommend using the WS-Trust bindings that ship with WIF to talk to ADFS 2 – they are pre-configured to match the binding configuration of the ADFS 2 endpoints. The following code requests a token for a WCF service from ADFS 2: private static SecurityToken GetToken() {     // Windows authentication over transport security     var factory = new WSTrustChannelFactory(         new WindowsWSTrustBinding(SecurityMode.Transport),         stsEndpoint);     factory.TrustVersion = TrustVersion.WSTrust13;       var rst = new RequestSecurityToken     {         RequestType = RequestTypes.Issue,         AppliesTo = new EndpointAddress(svcEndpoint),         KeyType = KeyTypes.Symmetric     };       var channel = factory.CreateChannel();     return channel.Issue(rst); } Afterwards, the returned token can be used to create a channel to the service. Again WIF has some helper methods here that make this very easy: private static void CallService(SecurityToken token) {     // create binding and turn off sessions     var binding = new WS2007FederationHttpBinding(         WSFederationHttpSecurityMode.TransportWithMessageCredential);     binding.Security.Message.EstablishSecurityContext = false;       // create factory and enable WIF plumbing     var factory = new ChannelFactory<IService>(binding, new EndpointAddress(svcEndpoint));     factory.ConfigureChannelFactory<IService>();       // turn off CardSpace - we already have the token     factory.Credentials.SupportInteractive = false;       var channel = factory.CreateChannelWithIssuedToken<IService>(token);       channel.GetClaims().ForEach(c =>         Console.WriteLine("{0}\n {1}\n  {2} ({3})\n",             c.ClaimType,             c.Value,             c.Issuer,             c.OriginalIssuer)); } Why is this approach more flexible? Well – some don’t like the configuration voodoo. That’s a valid reason for using the manual approach. You also get more control over the token request itself since you have full control over the RST message that gets send to the STS. One common parameter that you may want to set yourself is the appliesTo value. When you use the automatic token support in the WCF federation binding, the appliesTo is always the physical service address. This means in turn that this address will be used as the audience URI value in the SAML token. Well – this in turn means that when you have an application that consists of multiple services, you always have to configure all physical endpoint URLs in ADFS 2 and in the WIF configuration of the service(s). Having control over the appliesTo allows you to use more symbolic realm names, e.g. the base address or a completely logical name. Since the URL is never de-referenced you have some degree of freedom here. In the next post we will look at the necessary code to request multiple tokens in a call chain. This is a common scenario when you first have to acquire a token from an identity provider and have to send that on to a federation gateway or Resource STS. Stay tuned.

    Read the article

  • TFS API Add Favorites programmatically

    - by Tarun Arora
    01 – What are we trying to achieve? In this blog post I’ll be showing you how to add work item queries as favorites, it is also possible to use the same technique to add build definition as favorites. Once a shared query or build definition has been added as favorite it will show up on the team web access.  In this blog post I’ll be showing you a work around in the absence of a proper API how you can add queries to team favorites. 02 – Disclaimer There is no official API for adding favorites programmatically. In the work around below I am using the Identity service to store this data in a property bag which is used during display of favorites on the team web site. This uses an internal data structure that could change over time, there is no guarantee about the key names or content of the values. What is shown below is a workaround for a missing API. 03 – Concept There is no direct API support for favorites, but you could work around it using the identity service in TFS.  Favorites are stored in the property bag associated with the TeamFoundationIdentity (either the ‘team’ identity or the users identity depending on if these are ‘team’ or ‘my’ favorites).  The data is stored as json in the property bag of the identity, the key being prefixed by ‘Microsoft.TeamFoundation.Framework.Server.IdentityFavorites’. References - Microsoft.TeamFoundation.WorkItemTracking.Client - using Microsoft.TeamFoundation.Client; - using Microsoft.TeamFoundation.Framework.Client; - using Microsoft.TeamFoundation.Framework.Common; - using Microsoft.TeamFoundation.ProcessConfiguration.Client; - using Microsoft.TeamFoundation.Server; - using Microsoft.TeamFoundation.WorkItemTracking.Client; Services - IIdentityManagementService2 - TfsTeamService - WorkItemStore 04 – Solution Lets start by connecting to TFS programmatically // Create an instance of the services to be used during the program private static TfsTeamProjectCollection _tfs; private static ProjectInfo _selectedTeamProject; private static WorkItemStore _wis; private static TfsTeamService _tts; private static TeamSettingsConfigurationService _teamConfig; private static IIdentityManagementService2 _ids; // Connect to TFS programmatically public static bool ConnectToTfs() { var isSelected = false; var tfsPp = new TeamProjectPicker(TeamProjectPickerMode.SingleProject, false); tfsPp.ShowDialog(); _tfs = tfsPp.SelectedTeamProjectCollection; if (tfsPp.SelectedProjects.Any()) { _selectedTeamProject = tfsPp.SelectedProjects[0]; isSelected = true; } return isSelected; } Lets get all the work item queries from the selected team project static readonly Dictionary<string, string> QueryAndGuid = new Dictionary<string, string>(); // Get all queries and query guid in the selected team project private static void GetQueryGuidList(IEnumerable<QueryItem> query) { foreach (QueryItem subQuery in query) { if (subQuery.GetType() == typeof(QueryFolder)) GetQueryGuidList((QueryFolder)subQuery); else { QueryAndGuid.Add(subQuery.Name, subQuery.Id.ToString()); } } }   Pass the name of a valid Team in your team project and a name of a valid query in your team project. The team details will be extracted using the team name and query GUID will be extracted using the query name. These details will be used to construct the key and value that will be passed to the SetProperty method in the Identity service.           Key           “Microsoft.TeamFoundation.Framework.Server.IdentityFavorites..<TeamProjectURI>.<TeamId>.WorkItemTracking.Queries.<newGuid1>”           Value           "{"data":"<QueryGuid>","id":"<NewGuid1>","name":"<QueryKey>","type":"Microsoft.TeamFoundation.WorkItemTracking.QueryItem”}"           // Configure a Work Item Query for the given team private static void ConfigureTeamFavorites(string teamName, string queryName) { _ids = _tfs.GetService<IIdentityManagementService2>(); var g = Guid.NewGuid(); var guid = string.Empty; var teamDetail = _tts.QueryTeams(_selectedTeamProject.Uri).FirstOrDefault(t => t.Name == teamName); foreach (var q in QueryAndGuid.Where(q => q.Key == queryName)) { guid = q.Value; } if(guid == string.Empty) { Console.WriteLine("Query '{0}' - Not found!", queryName); return; } var key = string.Format( "Microsoft.TeamFoundation.Framework.Server.IdentityFavorites..{0}.{1}.WorkItemTracking.Queries{2}", new Uri(_selectedTeamProject.Uri).Segments.LastOrDefault(), teamDetail.Identity.TeamFoundationId, g); var value = string.Format( @"{0}""data"":""{1}"",""id"":""{2}"",""name"":""{3}"",""type"":""Microsoft.TeamFoundation.WorkItemTracking.QueryItem""{4}", "{", guid, g, QueryAndGuid.FirstOrDefault(q => q.Value==guid).Key, "}"); teamDetail.Identity.SetProperty(IdentityPropertyScope.Local, key, value); _ids.UpdateExtendedProperties(teamDetail.Identity); Console.WriteLine("{0}Added Query '{1}' as Favorite", Environment.NewLine, queryName); }   If you have any questions or suggestions leave a comment. Enjoy!

    Read the article

  • Automating custom software installation in a zone

    - by mgerdts
    In Solaris 11, the internals of zone installation are quite different than they were in Solaris 10.  This difference allows the administrator far greater control of what software is installed in a zone.  The rules in Solaris 10 are simple and inflexible: if it is installed in the global zone and is not specifically excluded by package metadata from being installed in a zone, it is installed in the zone.  In Solaris 11, the rules are still simple, but are much more flexible:  the packages you tell it to install and the packages on which they depend will be installed. So, where does the default list of packages come from?  From the AI (auto installer) manifest, of course.  The default AI manifest is /usr/share/auto_install/manifest/zone_default.xml.  Within that file you will find:             <software_data action="install">                 <name>pkg:/group/system/solaris-small-server</name>             </software_data> So, the default installation will install pkg:/group/system/solaris-small-server.  Cool.  What is that?  You can figure out what is in the package by looking for it in the repository with your web browser (click the manifest link), or use pkg(1).  In this case, it is a group package (pkg:/group/), so we know that it just has a bunch of dependencies to name the packages that really wants installed. $ pkg contents -t depend -o fmri -s fmri -r solaris-small-server FMRI compress/bzip2 compress/gzip compress/p7zip ... terminal/luit terminal/resize text/doctools text/doctools/ja text/less text/spelling-utilities web/wget If you would like to see the entire manifest from the command line, use pkg contents -r -m solaris-small-server. Let's suppose that you want to install a zone that also has mercurial and a full-fledged installation of vim rather than just the minimal vim-core that is part of solaris-small-server.  That's pretty easy. First, copy the default AI manifest somewhere where you will edit it and make it writable. # cp /usr/share/auto_install/manifest/zone_default.xml ~/myzone-ai.xml # chmod 644 ~/myzone-ai.xml Next, edit the file, changing the software_data section as follows:             <software_data action="install">                 <name>pkg:/group/system/solaris-small-server</name>                 <name>pkg:/developer/versioning/mercurial</name>                <name>pkg:/editor/vim</name>             </software_data> To figure out  the names of the packages, either search the repository using your browser, or use a command like pkg search hg. Now we are all ready to install the zone.  If it has not yet been configured, that must be done as well. # zonecfg -z myzone 'create; set zonepath=/zones/myzone' # zoneadm -z myzone install -m ~/myzone-ai.xml A ZFS file system has been created for this zone. Progress being logged to /var/log/zones/zoneadm.20111113T004303Z.myzone.install Image: Preparing at /zones/myzone/root. Install Log: /system/volatile/install.15496/install_log AI Manifest: /tmp/manifest.xml.XfaWpE SC Profile: /usr/share/auto_install/sc_profiles/enable_sci.xml Zonename: myzone Installation: Starting ... Creating IPS image Installing packages from: solaris origin: http://localhost:1008/solaris/54453f3545de891d4daa841ddb3c844fe8804f55/ DOWNLOAD PKGS FILES XFER (MB) Completed 169/169 34047/34047 185.6/185.6 PHASE ACTIONS Install Phase 46498/46498 PHASE ITEMS Package State Update Phase 169/169 Image State Update Phase 2/2 Installation: Succeeded Note: Man pages can be obtained by installing pkg:/system/manual done. Done: Installation completed in 531.813 seconds. Next Steps: Boot the zone, then log into the zone console (zlogin -C) to complete the configuration process. Log saved in non-global zone as /zones/myzone/root/var/log/zones/zoneadm.20111113T004303Z.myzone.install Now, for a few things that I've seen people trip over: Ignore that bit about man pages - it's wrong.  Man pages are already installed so long as the right facet is set properly.  And that's a topic for another blog entry. If you boot the zone then just use zlogin myzone, you will see that services you care about haven't started and that svc:/milestone/config:default is starting.  That is because you have not yet logged into the console with zlogin -C myzone. If the zone has been booted for more than a very short while when you first connect to the zone console, it will seem like the console is hung.  That's not really the case - hit ^L (control-L) to refresh the sysconfig(1M) screen that is prompting you for information.

    Read the article

  • 2D Platformer Collision Handling

    - by defender-zone
    Hello, everyone! I am trying to create a 2D platformer (Mario-type) game and I am some having some issues with handling collisions properly. I am writing this game in C++, using SDL for input, image loading, font loading, etcetera. I am also using OpenGL via the FreeGLUT library in conjunction with SDL to display graphics. My method of collision detection is AABB (Axis-Aligned Bounding Box), which is really all I need to start with. What I need is an easy way to both detect which side the collision occurred on and handle the collisions properly. So, basically, if the player collides with the top of the platform, reposition him to the top; if there is a collision to the sides, reposition the player back to the side of the object; if there is a collision to the bottom, reposition the player under the platform. I have tried many different ways of doing this, such as trying to find the penetration depth and repositioning the player backwards by the penetration depth. Sadly, nothing I've tried seems to work correctly. Player movement ends up being very glitchy and repositions the player when I don't want it to. Part of the reason is probably because I feel like this is something so simple but I'm over-thinking it. If anyone thinks they can help, please take a look at the code below and help me try to improve on this if you can. I would like to refrain from using a library to handle this (as I want to learn on my own) or the something like the SAT (Separating Axis Theorem) if at all possible. Thank you in advance for your help! void world1Level1CollisionDetection() { for(int i; i < blocks; i++) { if (de2dCheckCollision(ball,block[i],0.0f,0.0f)==true) { int up = 0; int left = 0; int right = 0; int down = 0; if(ball.coords[0] < block[i].coords[0] && block[i].coords[0] < ball.coords[2] && ball.coords[2] < block[i].coords[2]) { left = 1; } if(block[i].coords[0] < ball.coords[0] && ball.coords[0] < block[i].coords[2] && block[i].coords[2] < ball.coords[2]) { right = 1; } if(ball.coords[1] < block[i].coords[1] && block[i].coords[1] < ball.coords[3] && ball.coords[3] < block[i].coords[3]) { up = 1; } if(block[i].coords[1] < ball.coords[1] && ball.coords[1] < block[i].coords[3] && block[i].coords[3] < ball.coords[3]) { down = 1; } cout << left << ", " << right << ", " << up << ", " << down << ", " << endl; if (left == 1) { ball.coords[0] = block[i].coords[0] - 16.0f; ball.coords[2] = block[i].coords[0] - 0.0f; } if (right == 1) { ball.coords[0] = block[i].coords[2] + 0.0f; ball.coords[2] = block[i].coords[2] + 16.0f; } if (down == 1) { ball.coords[1] = block[i].coords[3] + 0.0f; ball.coords[3] = block[i].coords[3] + 16.0f; } if (up == 1) { ball.yspeed = 0.0f; ball.gravity = 0.0f; ball.coords[1] = block[i].coords[1] - 16.0f; ball.coords[3] = block[i].coords[1] - 0.0f; } } if (de2dCheckCollision(ball,block[i],0.0f,0.0f)==false) { ball.gravity = -0.5f; } } } To explain what some of this code means: The blocks variable is basically an integer that is storing the amount of blocks, or platforms. I am checking all of the blocks using a for loop, and the number that the loop is currently on is represented by integer i. The coordinate system might seem a little weird, so that's worth explaining. coords[0] represents the x position (left) of the object (where it starts on the x axis). coords[1] represents the y position (top) of the object (where it starts on the y axis). coords[2] represents the width of the object plus coords[0] (right). coords[3] represents the height of the object plus coords[1] (bottom). de2dCheckCollision performs an AABB collision detection. Up is negative y and down is positive y, as it is in most games. Hopefully I have provided enough information for someone to help me successfully. If there is something I left out that might be crucial, let me know and I'll provide the necessary information. Finally, for anyone who can help, providing code would be very helpful and much appreciated. Thank you again for your help!

    Read the article

  • BPM 11g - Dynamic Task Assignment with Multi-level Organization Units

    - by Mark Foster
    I've seen several requirements to have a more granular level of task assignment in BPM 11g based on some value in the data passed to the process. Parametric Roles is normally the first port of call to try to satisfy this requirement, but in this blog we will show how a lot of use-cases can be satisfied by the easier to implement and flexible Organization Unit. The Use-Case Task assignment is to an approval group containing several users. At runtime, a location value in the input data determines which of the particular users the task is ultimately assigned to. In this case we use the Demo Community referenced in the SOA Admin Guide, and specifically the "LoanAnalyticGroup" which contains three users; "szweig", "mmitch" & "fkafka". In our scenario we would like to assign a task to "szweig" if the input data specifies that the location is "JapanCentral", to "fkafka" if the location is "JapanNorth" and to "mmitch" if "JapanSouth", and to all of them if the location is "Japan" i.e....   The Process Simple one human task process.... In the output data association of the "Start" activity we need to set the value of the "Organization Unit" predefined variable based on the input data (note that the  predefined variables can only be set on output data associations)....  ...and in the output data association of the human activity we will reset the "Organization Unit" to empty, always good practice to ensure that the Organization Unit will not be used for any subsequent human activities for which we do not require it.... Set Up the Organization Unit  Log in to the BPM Workspace with an administrator user (weblogic/welcome1 in our case) and choose the "Administration" option. Within "Roles" assign the "ProcessOwner" swim-lane for our process to "LoanAnalyticGroup".... Within "Organization Units" we can model our organization.... "Root Organization Unit" as "Japan" and "Child Organization Unit" as "Central", "South" & "North" as shown. As described previously, add user "szweig" to "Central", "mmitch" to "South" and "fkafka" to "North"....   Test the Process Invalid Data  First let us test with invalid data in the input to see what the consequences are, here we use "X" as input.... ...and looking at the instance we can see it has errored.... Organization Unit Root Level Assignment  Now let us see what happens if we have "Japan" in the input data.... ...looking in the "flow trace" we can see that the task has been assigned....  ... but who has the task been assigned to ? Let us look in the BPM Workspace for user "szweig"....  ...and for "mmitch"....  ... and for "fkafka"....  ...so we can see that with an Organization Unit at "Root" level we have successfully assigned the task to all users. Organization Unit Child Level Assignment  Now let us test with "Japan/North" in the input data.... ...and looking in "fkafka" workspace we see the task has been assigned, remember, he was associated with "JapanNorth"....   ... but what about the workspace of "szweig"....  ...no tasks assigned, neither has "mmitch", just as we expected. Summary  We have seen in this blog how to easily implement multi-level dynamic task routing using Organization Units, a common use-case and a simpler solution than Parametric Roles. 

    Read the article

  • Scripting Part 1

    - by rbishop
    Dynamic Scripting is a large topic, so let me get a couple of things out of the way first. If you aren't familiar with JavaScript, I can suggest CodeAcademy's JavaScript series. There are also many other websites and books that cover JavaScript from every possible angle.The second thing we need to deal with is JavaScript as a programming language versus a JavaScript environment running in a web browser. Many books, tutorials, and websites completely blur these two together but they are in fact completely separate. What does this really mean in relation to DRM? Since DRM isn't a web browser, there are no document, window, history, screen, or location objects. There are no events like mousedown or click. Trying to call alert('hello!') in DRM will just cause an error. Those concepts are all related to an HTML document (web page) and are part of the Browser Object Model or Document Object Model. DRM has its own object model that exposes DRM-related objects. In practice, feel free to use those sorts of tutorials or practice within your browser; Many of the concepts are directly translatable to writing scripts in DRM. Just don't try to call document.getElementById in your property definition!I think learning by example tends to work the best, so let's try getting a list of all the unique property values for a given node and its children. var uniqueValues = {}; var childEnumerator = node.GetChildEnumerator(); while(childEnumerator.MoveNext()) { var propValue = childEnumerator.GetCurrent().PropValue("Custom.testpropstr1"); print(propValue); if(propValue != null && propValue != '' && !uniqueValues[propValue]) uniqueValues[propValue] = true; } var result = ''; for(var value in uniqueValues){ result += "Found value " + value + ","; } return result;  Now lets break this down piece by piece. var uniqueValues = {}; This declares a variable and initializes it as a new empty Object. You could also have written var uniqueValues = new Object(); Why use an object here? JavaScript objects can also function as a list of keys and we'll use that later to store each property value as a key on the object. var childEnumerator = node.GetChildEnumerator(); while(childEnumerator.MoveNext()) { This gets an enumerator for the node's children. The enumerator allows us to loop through the children one by one. If we wanted to get a filtered list of children, we would instead use ChildrenWith(). When we reach the end of the child list, the enumerator will return false for MoveNext() and that will stop the loop. var propValue = childEnumerator.GetCurrent().PropValue("Custom.testpropstr1"); print(propValue); if(propValue != null && propValue != '' && !uniqueValues[propValue]) uniqueValues[propValue] = true; } This gets the node the enumerator is currently pointing at, then calls PropValue() on it to get the value of a property. We then make sure the prop value isn't null or the empty string, then we make sure the value doesn't already exist as a key. Assuming it doesn't we add it as a key with a value (true in this case because it makes checking for an existing value faster when the value exists). A quick word on the print() function. When viewing the prop grid, running an export, or performing normal DRM operations it does nothing. If you have a lot of print() calls with complicated arguments it can slow your script down slightly, but otherwise has no effect. But when using the script editor, all the output of print() will be shown in the Warnings area. This gives you an extremely useful debugging tool to see what exactly a script is doing. var result = ''; for(var value in uniqueValues){ result += "Found value " + value + ","; } return result; Now we build a string by looping through all the keys in uniqueValues and adding that value to our string. The last step is to simply return the result. Hopefully this small example demonstrates some of the core Dynamic Scripting concepts. Next time, we can try checking for node references in other hierarchies to see if they are using duplicate property values.

    Read the article

  • Converting LINQ to Twitter to Twitter API v1.1

    - by Joe Mayo
    Twitter recently updated their API to v1.1 (Current status: API v1.1). Naturally, LINQ to Twitter  needed to be updated too. This blog post outlines the changes made to LINQ to Twitter during this conversion and highlights important features that LINQ to Twitter developers will want to know. Overall Impact Generally speaking, Twitter API v1.1 is semantically very much the same as it’s predecessor. The base URL changed and so did a few resource segments, but the resources themselves are still intact. The good news is that LINQ to Twitter has always shielded the developer from this plumbing, so the entities, types, and filters didn’t change much at all.  The following sections describe what did  change. Authentication In Twitter API v1.0 authentication was not required for some resources, such as user timelines and search. However, that’s all changed because *all* queries must be authenticated in Twitter API v1.1. LINQ to Twitter has various types of authorizers you can use, supporting whatever OAuth options are available via Twitter.  You can see the LINQ to Twitter documentation, Securing Your Applications, for more info on OAuth support. The New Search One of the larger changes to the API was Search. To be more specific, the Search entity now contains a List<Status>, named Statuses, to hold results.  Additionally, any meta-data associated with the search is now in a property named SearchMetaData. The change to the Search entity and responses is the big change, but the good news is that your Search query syntax doesn’t change. Different Rate Limits The issue of rate limits itself is contentious, but this discussion is focused on the coding experience and I’ll leave the politics to those who prefer to engage in that activity. What’s important here is that both headers and resources have changed. You should review Twitter’s Rate Limit documentation to understand what the changes mean.  A quick explanation is that rate limits are applied individually to each resource in 15 minute time intervals. In LINQ to Twitter these changes surface on the Help entity, via HelpType.RateLimits. The RateLimits query has a Resources filter where you can specify a comma-separated list of categories to return rate limit info for.  The results materialize in the RateLimits dictionary, keyed on category. The Help entity also has a RateLimitsAuthorizationContext, holding the Access Token for the user performing queries – and to whom the rate limits apply. In addition to the new RateLimits query, there are new RateLimit headers that appear in the query response, whose HTTP header name is of the form X-Rate-Limit… which is different from the previous header name. LINQ to Twitter surfaces these headers via the existing properties of the TwitterContext instance. For anyone who retrieved rate limit information via the Headers property of TwitterContext, you should be aware of the new header names.  I haven’t done anything with Feature rate limit properties yet, but they appear to no longer be available – this will require more follow-up. Error Handling Twitter API v1.1 has a new format for Error Codes & Responses. LINQ to Twitter wraps these messages in the TwitterQueryException, which has been updated appropriately. The Message property of TwitterQueryException now reflects the Twitter error message, when available. There’s also a new ErrorCode that’s populated with the message error code. Parameters Most parameters stayed the same, but one of interest is Include Entities (different from LINQ to Twitter data object entities). Entities are metadata hanging off tweets, that provide start/end position in the tweet and other information for mentions, urls, hash tags, and media. Entities used to not be included unless you specified you wanted them. Now, in v1.1, entities are included by default for all APIs that return a Status.  If you were always setting IncludeEntities to true, then you won’t see a change. However, be aware that you’ll now be receiving additional data in your response from Twitter, which will explain a sudden increase in bandwidth utilization. This might or might not  matter to you  depending on the requirements of your application, but you should be aware of it. Everything Else There might be small changes here and there that I haven’t mentioned, but these were the ones you should be most aware of.  Streams didn’t change, but Twitter will be deprecating username/password authentication on public streams, in favor of OAuth, so you’ll be seeing me make that change some time in the future.  Also, Twitter will continue to evolve the API and you can expect that LINQ to Twitter will change accordingly. Summary The big changes to Twitter API were Authentication, Search, Rate Limits, and Error Handling. All API calls must be authenticated. You’ll need to change your code to read Search results differently, but the query is much the same as you use now. There’s a new RateLimits API, one of the Help queries.  Also, the new error messages are integrated into TwitterQueryException. Besides these changes, I expect  most others to be small or affect a smaller percentage of developers.  You can get the latest version of LINQ to Twitter from NuGet or visit the LINQ to Twitter download page at CodePlex.com.   @JoeMayo

    Read the article

  • Mapping Repeating Sequence Groups in BizTalk

    - by Paul Petrov
    Repeating sequence groups can often be seen in real life XML documents. It happens when certain sequence of elements repeats in the instance document. Here’s fairly abstract example of schema definition that contains sequence group: <xs:schemaxmlns:b="http://schemas.microsoft.com/BizTalk/2003"            xmlns:xs="http://www.w3.org/2001/XMLSchema"            xmlns="NS-Schema1"            targetNamespace="NS-Schema1" >  <xs:elementname="RepeatingSequenceGroups">     <xs:complexType>       <xs:sequencemaxOccurs="1"minOccurs="0">         <xs:sequencemaxOccurs="unbounded">           <xs:elementname="A"type="xs:string" />           <xs:elementname="B"type="xs:string" />           <xs:elementname="C"type="xs:string"minOccurs="0" />         </xs:sequence>       </xs:sequence>     </xs:complexType>  </xs:element> </xs:schema> And here’s corresponding XML instance document: <ns0:RepeatingSequenceGroupsxmlns:ns0="NS-Schema1">  <A>A1</A>  <B>B1</B>  <C>C1</C>  <A>A2</A>  <B>B2</B>  <A>A3</A>  <B>B3</B>  <C>C3</C> </ns0:RepeatingSequenceGroups> As you can see elements A, B, and C are children of anonymous xs:sequence element which in turn can be repeated N times. Let’s say we need do simple mapping to the schema with similar structure but with different element names: <ns0:Destinationxmlns:ns0="NS-Schema2">  <Alpha>A1</Alpha>  <Beta>B1</Beta>  <Gamma>C1</Gamma>  <Alpha>A2</Alpha>  <Beta>B2</Beta>  <Gamma>C2</Gamma> </ns0:Destination> The basic map for such typical task would look pretty straightforward: If we test this map without any modification it will produce following result: <ns0:Destinationxmlns:ns0="NS-Schema2">  <Alpha>A1</Alpha>  <Alpha>A2</Alpha>  <Alpha>A3</Alpha>  <Beta>B1</Beta>  <Beta>B2</Beta>  <Beta>B3</Beta>  <Gamma>C1</Gamma>  <Gamma>C3</Gamma> </ns0:Destination> The original order of the elements inside sequence is lost and that’s not what we want. Default behavior of the BizTalk 2009 and 2010 Map Editor is to generate compatible map with older versions that did not have ability to preserve sequence order. To enable this feature simply open map file (*.btm) in text/xml editor and find attribute PreserveSequenceOrder of the root <mapsource> element. Set its value to Yes and re-test the map: <ns0:Destinationxmlns:ns0="NS-Schema2">  <Alpha>A1</Alpha>  <Beta>B1</Beta>  <Gamma>C1</Gamma>  <Alpha>A2</Alpha>  <Beta>B2</Beta>  <Alpha>A3</Alpha>  <Beta>B3</Beta>  <Gamma>C3</Gamma> </ns0:Destination> The result is as expected – all corresponding elements are in the same order as in the source document. Under the hood it is achieved by using one common xsl:for-each statement that pulls all elements in original order (rather than using individual for-each statement per element name in default mode) and xsl:if statements to test current element in the loop:  <xsl:templatematch="/s0:RepeatingSequenceGroups">     <ns0:Destination>       <xsl:for-eachselect="A|B|C">         <xsl:iftest="local-name()='A'">           <Alpha>             <xsl:value-ofselect="./text()" />           </Alpha>         </xsl:if>         <xsl:iftest="local-name()='B'">           <Beta>             <xsl:value-ofselect="./text()" />           </Beta>         </xsl:if>         <xsl:iftest="local-name()='C'">           <Gamma>             <xsl:value-ofselect="./text()" />           </Gamma>         </xsl:if>       </xsl:for-each>     </ns0:Destination>  </xsl:template> BizTalk Map editor became smarter so learn and use this lesser known feature of XSLT 2.0 in your maps and XSL stylesheets.

    Read the article

  • Rkhunter 122 suspect files; do I have a problem?

    - by user276166
    I am new to ubuntu. I am using Xfce Ubuntu 14.04 LTS. I have ran rkhunter a few weeks age and only got a few warnings. The forum said that they were normal. But, this time rkhunter reported 122 warnings. Please advise. casey@Shaman:~$ sudo rkhunter -c [ Rootkit Hunter version 1.4.0 ] Checking system commands... Performing 'strings' command checks Checking 'strings' command [ OK ] Performing 'shared libraries' checks Checking for preloading variables [ None found ] Checking for preloaded libraries [ None found ] Checking LD_LIBRARY_PATH variable [ Not found ] Performing file properties checks Checking for prerequisites [ Warning ] /usr/sbin/adduser [ Warning ] /usr/sbin/chroot [ Warning ] /usr/sbin/cron [ OK ] /usr/sbin/groupadd [ Warning ] /usr/sbin/groupdel [ Warning ] /usr/sbin/groupmod [ Warning ] /usr/sbin/grpck [ Warning ] /usr/sbin/nologin [ Warning ] /usr/sbin/pwck [ Warning ] /usr/sbin/rsyslogd [ Warning ] /usr/sbin/useradd [ Warning ] /usr/sbin/userdel [ Warning ] /usr/sbin/usermod [ Warning ] /usr/sbin/vipw [ Warning ] /usr/bin/awk [ Warning ] /usr/bin/basename [ Warning ] /usr/bin/chattr [ Warning ] /usr/bin/cut [ Warning ] /usr/bin/diff [ Warning ] /usr/bin/dirname [ Warning ] /usr/bin/dpkg [ Warning ] /usr/bin/dpkg-query [ Warning ] /usr/bin/du [ Warning ] /usr/bin/env [ Warning ] /usr/bin/file [ Warning ] /usr/bin/find [ Warning ] /usr/bin/GET [ Warning ] /usr/bin/groups [ Warning ] /usr/bin/head [ Warning ] /usr/bin/id [ Warning ] /usr/bin/killall [ OK ] /usr/bin/last [ Warning ] /usr/bin/lastlog [ Warning ] /usr/bin/ldd [ Warning ] /usr/bin/less [ OK ] /usr/bin/locate [ OK ] /usr/bin/logger [ Warning ] /usr/bin/lsattr [ Warning ] /usr/bin/lsof [ OK ] /usr/bin/mail [ OK ] /usr/bin/md5sum [ Warning ] /usr/bin/mlocate [ OK ] /usr/bin/newgrp [ Warning ] /usr/bin/passwd [ Warning ] /usr/bin/perl [ Warning ] /usr/bin/pgrep [ Warning ] /usr/bin/pkill [ Warning ] /usr/bin/pstree [ OK ] /usr/bin/rkhunter [ OK ] /usr/bin/rpm [ Warning ] /usr/bin/runcon [ Warning ] /usr/bin/sha1sum [ Warning ] /usr/bin/sha224sum [ Warning ] /usr/bin/sha256sum [ Warning ] /usr/bin/sha384sum [ Warning ] /usr/bin/sha512sum [ Warning ] /usr/bin/size [ Warning ] /usr/bin/sort [ Warning ] /usr/bin/stat [ Warning ] /usr/bin/strace [ Warning ] /usr/bin/strings [ Warning ] /usr/bin/sudo [ Warning ] /usr/bin/tail [ Warning ] /usr/bin/test [ Warning ] /usr/bin/top [ Warning ] /usr/bin/touch [ Warning ] /usr/bin/tr [ Warning ] /usr/bin/uniq [ Warning ] /usr/bin/users [ Warning ] /usr/bin/vmstat [ Warning ] /usr/bin/w [ Warning ] /usr/bin/watch [ Warning ] /usr/bin/wc [ Warning ] /usr/bin/wget [ Warning ] /usr/bin/whatis [ Warning ] /usr/bin/whereis [ Warning ] /usr/bin/which [ OK ] /usr/bin/who [ Warning ] /usr/bin/whoami [ Warning ] /usr/bin/unhide.rb [ Warning ] /usr/bin/mawk [ Warning ] /usr/bin/lwp-request [ Warning ] /usr/bin/heirloom-mailx [ OK ] /usr/bin/w.procps [ Warning ] /sbin/depmod [ Warning ] /sbin/fsck [ Warning ] /sbin/ifconfig [ Warning ] /sbin/ifdown [ Warning ] /sbin/ifup [ Warning ] /sbin/init [ Warning ] /sbin/insmod [ Warning ] /sbin/ip [ Warning ] /sbin/lsmod [ Warning ] /sbin/modinfo [ Warning ] /sbin/modprobe [ Warning ] /sbin/rmmod [ Warning ] /sbin/route [ Warning ] /sbin/runlevel [ Warning ] /sbin/sulogin [ Warning ] /sbin/sysctl [ Warning ] /bin/bash [ Warning ] /bin/cat [ Warning ] /bin/chmod [ Warning ] /bin/chown [ Warning ] /bin/cp [ Warning ] /bin/date [ Warning ] /bin/df [ Warning ] /bin/dmesg [ Warning ] /bin/echo [ Warning ] /bin/ed [ OK ] /bin/egrep [ Warning ] /bin/fgrep [ Warning ] /bin/fuser [ OK ] /bin/grep [ Warning ] /bin/ip [ Warning ] /bin/kill [ Warning ] /bin/less [ OK ] /bin/login [ Warning ] /bin/ls [ Warning ] /bin/lsmod [ Warning ] /bin/mktemp [ Warning ] /bin/more [ Warning ] /bin/mount [ Warning ] /bin/mv [ Warning ] /bin/netstat [ Warning ] /bin/ping [ Warning ] /bin/ps [ Warning ] /bin/pwd [ Warning ] /bin/readlink [ Warning ] /bin/sed [ Warning ] /bin/sh [ Warning ] /bin/su [ Warning ] /bin/touch [ Warning ] /bin/uname [ Warning ] /bin/which [ OK ] /bin/kmod [ Warning ] /bin/dash [ Warning ] [Press <ENTER> to continue] Checking for rootkits... Performing check of known rootkit files and directories 55808 Trojan - Variant A [ Not found ] ADM Worm [ Not found ] AjaKit Rootkit [ Not found ] Adore Rootkit [ Not found ] aPa Kit [ Not found ] Apache Worm [ Not found ] Ambient (ark) Rootkit [ Not found ] Balaur Rootkit [ Not found ] BeastKit Rootkit [ Not found ] beX2 Rootkit [ Not found ] BOBKit Rootkit [ Not found ] cb Rootkit [ Not found ] CiNIK Worm (Slapper.B variant) [ Not found ] Danny-Boy's Abuse Kit [ Not found ] Devil RootKit [ Not found ] Dica-Kit Rootkit [ Not found ] Dreams Rootkit [ Not found ] Duarawkz Rootkit [ Not found ] Enye LKM [ Not found ] Flea Linux Rootkit [ Not found ] Fu Rootkit [ Not found ] Fuck`it Rootkit [ Not found ] GasKit Rootkit [ Not found ] Heroin LKM [ Not found ] HjC Kit [ Not found ] ignoKit Rootkit [ Not found ] IntoXonia-NG Rootkit [ Not found ] Irix Rootkit [ Not found ] Jynx Rootkit [ Not found ] KBeast Rootkit [ Not found ] Kitko Rootkit [ Not found ] Knark Rootkit [ Not found ] ld-linuxv.so Rootkit [ Not found ] Li0n Worm [ Not found ] Lockit / LJK2 Rootkit [ Not found ] Mood-NT Rootkit [ Not found ] MRK Rootkit [ Not found ] Ni0 Rootkit [ Not found ] Ohhara Rootkit [ Not found ] Optic Kit (Tux) Worm [ Not found ] Oz Rootkit [ Not found ] Phalanx Rootkit [ Not found ] Phalanx2 Rootkit [ Not found ] Phalanx2 Rootkit (extended tests) [ Not found ] Portacelo Rootkit [ Not found ] R3dstorm Toolkit [ Not found ] RH-Sharpe's Rootkit [ Not found ] RSHA's Rootkit [ Not found ] Scalper Worm [ Not found ] Sebek LKM [ Not found ] Shutdown Rootkit [ Not found ] SHV4 Rootkit [ Not found ] SHV5 Rootkit [ Not found ] Sin Rootkit [ Not found ] Slapper Worm [ Not found ] Sneakin Rootkit [ Not found ] 'Spanish' Rootkit [ Not found ] Suckit Rootkit [ Not found ] Superkit Rootkit [ Not found ] TBD (Telnet BackDoor) [ Not found ] TeLeKiT Rootkit [ Not found ] T0rn Rootkit [ Not found ] trNkit Rootkit [ Not found ] Trojanit Kit [ Not found ] Tuxtendo Rootkit [ Not found ] URK Rootkit [ Not found ] Vampire Rootkit [ Not found ] VcKit Rootkit [ Not found ] Volc Rootkit [ Not found ] Xzibit Rootkit [ Not found ] zaRwT.KiT Rootkit [ Not found ] ZK Rootkit [ Not found ] [Press <ENTER> to continue] Performing additional rootkit checks Suckit Rookit additional checks [ OK ] Checking for possible rootkit files and directories [ None found ] Checking for possible rootkit strings [ None found ] Performing malware checks Checking running processes for suspicious files [ None found ] Checking for login backdoors [ None found ] Checking for suspicious directories [ None found ] Checking for sniffer log files [ None found ] Performing Linux specific checks Checking loaded kernel modules [ OK ] Checking kernel module names [ OK ] [Press <ENTER> to continue] Checking the network... Performing checks on the network ports Checking for backdoor ports [ None found ] Checking for hidden ports [ Skipped ] Performing checks on the network interfaces Checking for promiscuous interfaces [ None found ] Checking the local host... Performing system boot checks Checking for local host name [ Found ] Checking for system startup files [ Found ] Checking system startup files for malware [ None found ] Performing group and account checks Checking for passwd file [ Found ] Checking for root equivalent (UID 0) accounts [ None found ] Checking for passwordless accounts [ None found ] Checking for passwd file changes [ Warning ] Checking for group file changes [ Warning ] Checking root account shell history files [ None found ] Performing system configuration file checks Checking for SSH configuration file [ Not found ] Checking for running syslog daemon [ Found ] Checking for syslog configuration file [ Found ] Checking if syslog remote logging is allowed [ Not allowed ] Performing filesystem checks Checking /dev for suspicious file types [ Warning ] Checking for hidden files and directories [ Warning ] [Press <ENTER> to continue] System checks summary ===================== File properties checks... Required commands check failed Files checked: 137 Suspect files: 122 Rootkit checks... Rootkits checked : 291 Possible rootkits: 0 Applications checks... All checks skipped The system checks took: 5 minutes and 11 seconds All results have been written to the log file (/var/log/rkhunter.log)

    Read the article

  • Disaster Recovery Discovery

    - by Rodney Landrum
    Last weekend I joined several of my IT staff on a mission to perform a DR test in our remote CoLo center in a large South East city of the US. Can I be more obtuse? The goal was simple for me as the sole DBA in a throng of Windows, Storage, Network and SAN admins – restore the databases and make them work. There were 4 applications that back ended to 7 SQL Server databases on 4 different SQL Server instances. We would maintain the original server names, but beyond that it was fair game. We had time to prepare so I was able to script out or otherwise automate the recovery process. I used sp_help_revlogin for three of the servers, a bit of a cheat actually because restoring the Master database on the target DR servers was the specified course of action according to the DR procedures ( the caveat “IF REQUIRED” left it open to interpretation. I really wanted to avoid the step of restoring Master for a number of reasons but mainly because I did not want to deal with issues starting SQL Services afterward. Having to account for the location of TempDB and the version conflicts of the resource DBs were just two of the battles I chose not to fight. Not to mention other system database location problems that might arise and prevent SQL from starting.  I was going to have to restore all of the user databases anyway, so I would not really gain any benefit, outside of logins, for taking the time to restore the source Master database over the newly installed one on the fresh server. What I wanted was the ability to restore the Master database as a user database, call it Master_Mine, from a backup on the source system and then use that restored database to script the SQL Logins and passwords on the DR systems. While I did not attempt this on the trip, the thought stuck in my mind and this past week I succeeded at scripting user accounts and passwords using only a restored copy of the Master database. Granted there were several challenges to overcome.  Also, as is usual for any work like this the usual disclaimers apply:  This is not something that I would imagine Microsoft would condone or support and this was really only an experiment for me to learn if it was even possible. While I have tested the process with success, I do not know that I would use this technique in a documented procedure because future updates for SQL Server will render this technique non-functional. I thought at first, incorrectly of course, that I could use sp_help_revlogin on a restored copy of the master database I named Master_Mine.   Since sp_help_revlogin uses system schema objects, sys.syslogins and sys.server_principals, this was not going to work because all results would come from the main Master database. To test this I added a SQL login via SSMS, backed up Master, restored  it as Master_Mine, and then deleted the login.  Even though the test account I created should presumably still be in the Master_Mine database, I should be able to get to it and script out its creation with its password hash so that I would not need to know the password, but any applications that stored that password would not have to be altered in the DR scenario. They would just work as expected. Once I realized that would not work I began looking deeper.  Knowing that sys.syslogins and sys.server_principals are system views, their underlying code should be available with sp_helptext, right? They were. And this led me to discover the two tables sys.sysxlgns and sys.sysprivs, where the data I needed was stored. These tables existed in both the real Master and the restored copy, Master_Mine.  I used this information to tweak the sp_help_revlogin stored procedure to use these tables instead to create the logins cursor used in sp_help_revlogin. For the password hash,  sp_help_revlogin uses the function LoginProperty() which takes a user name and option ‘passwordhash’ to return the hash for the user. Unfortunately, it requires the login to exist in the Master database. This would not work. So another slight modification I had to make was to pull the password hash itself (pwdhash from sys.sysxlgns) into the logins cursor and comment out the section of sp_help_revlogin that uses LoginProperty. Instead, I pass the pwdhash value as the variable @PWD_varbinary to the sp_hexadecimal stored procedure which is also created by and used within the code provided by Microsoft in the link above for sp_help_revlogin. The final challenge: sys.sysxlgns and sys.server_principals are visible only within a Dedicated Administrator Connection (DAC) query window in SSMS or within SQLCDMD.  To open a DAC connection you have to be logged in on the SQL Server itself, via RDP in my case,  and you preface the server name in the query connection with ADMIN:, so that the server connection looks like ADMIN:ServerName. From there you can create the modified stored procedure in the restored copy of a Master database from a source system as whatever name you like, and then run the modified stored procedure. I named my new stored procedure usp_help_revlogin_MyMaster. Upon execution I was happy to see the logins and password hashes that I needed to apply from the source Master database without having to restore over the new Master system database and without the need to access the original server (assuming it was down due to whatever disaster put it in that state). You will note that I am not providing full code samples here of the modifications. I will say that it was a slight bit of work and anyone who needed to do this for whatever reason, could fairly easily roll their own solution with the information provided herein.  My goal, as I said was to prove that this could be done and provide another option if required to ease the burden of getting SQL Servers up and available in an emergency situation where alternatives may be more challenging or otherwise unavailable.  

    Read the article

  • Physics System ignores collision in some rare cases

    - by Gajoo
    I've been developing a simple physics engine for my game. since the game physics is very simple I've decided to increase accuracy a little bit. Instead of formal integration methods like fourier or RK4, I'm directly computing the results after delta time "dt". based on the very first laws of physics : dx = 0.5 * a * dt^2 + v0 * dt dv = a * dt where a is acceleration and v0 is object's previous velocity. Also to handle collisions I've used a method which is somehow different from those I've seen so far. I'm detecting all the collision in the given time frame, stepping the world forward to the nearest collision, resolving it and again check for possible collisions. As I said the world consist of very simple objects, so I'm not loosing any performance due to multiple collision checking. First I'm checking if the ball collides with any walls around it (which is working perfectly) and then I'm checking if it collides with the edges of the walls (yellow points in the picture). the algorithm seems to work without any problem except some rare cases, in which the collision with points are ignored. I've tested everything and all the variables seem to be what they should but after leaving the system work for a minute or two the system the ball passes through one of those points. Here is collision portion of my code, hopefully one of you guys can give me a hint where to look for a potential bug! void PhysicalWorld::checkForPointCollision(Vec2 acceleration, PhysicsComponent& ball, Vec2& collisionNormal, float& collisionTime, Vec2 target) { // this function checks if there will be any collision between a circle and a point // ball contains informations about the circle (it's current velocity, position and radius) // collisionNormal is an output variable // collisionTime is also an output varialbe // target is the point I want to check for collisions Vec2 V = ball.mVelocity; Vec2 A = acceleration; Vec2 P = ball.mPosition - target; float wallWidth = mMap->getWallWidth() / (mMap->getWallWidth() + mMap->getHallWidth()) / 2; float r = ball.mRadius / (mMap->getWallWidth() + mMap->getHallWidth()); // r is ball radius scaled to match actual rendered object. if (A.any()) // todo : I need to first correctly solve the collisions in case there is no acceleration return; if (V.any()) // if object is not moving there will be no collisions! { float D = P.x * V.y - P.y * V.x; float Delta = r*r*V.length2() - D*D; if(Delta < eps) return; Delta = sqrt(Delta); float sgnvy = V.y > 0 ? 1: (V.y < 0?-1:0); Vec2 c1(( D*V.y+sgnvy*V.x*Delta) / V.length2(), (-D*V.x+fabs(V.y)*Delta) / V.length2()); Vec2 c2(( D*V.y-sgnvy*V.x*Delta) / V.length2(), (-D*V.x-fabs(V.y)*Delta) / V.length2()); float t1 = (c1.x - P.x) / V.x; float t2 = (c2.x - P.x) / V.x; if(t1 > eps && t1 <= collisionTime) { collisionTime = t1; collisionNormal = c1; } if(t2 > eps && t2 <= collisionTime) { collisionTime = t2; collisionNormal = c2; } } } // this function should step the world forward by dt. it doesn't check for collision of any two balls (components) // it just checks if there is a collision between the current component and 4 points forming a rectangle around it. void PhysicalWorld::step(float dt) { for (unsigned i=0;i<mObjects.size();i++) { PhysicsComponent &current = *mObjects[i]; Vec2 acceleration = current.mForces * current.mInvMass; float rt=dt; // stores how much more the world should advance while(rt > eps) { float collisionTime = rt; Vec2 collisionNormal = Vec2(0,0); float halfWallWidth = mMap->getWallWidth() / (mMap->getWallWidth() + mMap->getHallWidth()) / 2; // we check if there is any collision with any of those 4 points around the ball // if there is a collision both collisionNormal and collisionTime variables will change // after these functions collisionTime will be exactly the value of nearest collision (if any) // and if there was, collisionNormal will report in which direction the ball should return. checkForPointCollision(acceleration,current,collisionNormal,collisionTime,Vec2(floor(current.mPosition.x) + halfWallWidth,floor(current.mPosition.y) + halfWallWidth)); checkForPointCollision(acceleration,current,collisionNormal,collisionTime,Vec2(floor(current.mPosition.x) + halfWallWidth, ceil(current.mPosition.y) - halfWallWidth)); checkForPointCollision(acceleration,current,collisionNormal,collisionTime,Vec2( ceil(current.mPosition.x) - halfWallWidth,floor(current.mPosition.y) + halfWallWidth)); checkForPointCollision(acceleration,current,collisionNormal,collisionTime,Vec2( ceil(current.mPosition.x) - halfWallWidth, ceil(current.mPosition.y) - halfWallWidth)); // either if there is a collision or if there is not we step the forward since we are sure there will be no collision before collisionTime current.mPosition += collisionTime * (collisionTime * acceleration * 0.5 + current.mVelocity); current.mVelocity += collisionTime * acceleration; // if the ball collided with anything collisionNormal should be at least none zero in one of it's axis if (collisionNormal.any()) { collisionNormal *= Dot(collisionNormal, current.mVelocity) / collisionNormal.length2(); current.mVelocity -= 2 * collisionNormal; // simply reverse velocity along collision normal direction } rt -= collisionTime; } // reset all forces for current object so it'll be ready for later game event current.mForces.zero(); } }

    Read the article

  • WIF-less claim extraction from ACS: SWT

    - by Elton Stoneman
    WIF with SAML is solid and flexible, but unless you need the power, it can be overkill for simple claim assertion, and in the REST world WIF doesn’t have support for the latest token formats.  Simple Web Token (SWT) may not be around forever, but while it's here it's a nice easy format which you can manipulate in .NET without having to go down the WIF route. Assuming you have set up a Relying Party in ACS, specifying SWT as the token format: When ACS redirects to your login page, it will POST the SWT in the first form variable. It comes through in the BinarySecurityToken element of a RequestSecurityTokenResponse XML payload , the SWT type is specified with a TokenType of http://schemas.xmlsoap.org/ws/2009/11/swt-token-profile-1.0 : <t:RequestSecurityTokenResponse xmlns:t="http://schemas.xmlsoap.org/ws/2005/02/trust">   <t:Lifetime>     <wsu:Created xmlns:wsu="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd">2012-08-31T07:31:18.655Z</wsu:Created>     <wsu:Expires xmlns:wsu="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd">2012-08-31T09:11:18.655Z</wsu:Expires>   </t:Lifetime>   <wsp:AppliesTo xmlns:wsp="http://schemas.xmlsoap.org/ws/2004/09/policy">     <EndpointReference xmlns="http://www.w3.org/2005/08/addressing">       <Address>http://localhost/x.y.z</Address>     </EndpointReference>   </wsp:AppliesTo>   <t:RequestedSecurityToken>     <wsse:BinarySecurityToken wsu:Id="uuid:fc8d3332-d501-4bb0-84ba-d31aa95a1a6c" ValueType="http://schemas.xmlsoap.org/ws/2009/11/swt-token-profile-1.0" EncodingType="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-soap-message-security-1.0#Base64Binary" xmlns:wsu="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd" xmlns:wsse="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"> [ base64string ] </wsse:BinarySecurityToken>   </t:RequestedSecurityToken>   <t:TokenType>http://schemas.xmlsoap.org/ws/2009/11/swt-token-profile-1.0</t:TokenType>   <t:RequestType>http://schemas.xmlsoap.org/ws/2005/02/trust/Issue</t:RequestType>   <t:KeyType>http://schemas.xmlsoap.org/ws/2005/05/identity/NoProofKey</t:KeyType> </t:RequestSecurityTokenResponse> Reading the SWT is as simple as base-64 decoding, then URL-decoding the element value:     var wrappedToken = XDocument.Parse(HttpContext.Current.Request.Form[1]);     var binaryToken = wrappedToken.Root.Descendants("{http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd}BinarySecurityToken").First();     var tokenBytes = Convert.FromBase64String(binaryToken.Value);     var token = Encoding.UTF8.GetString(tokenBytes);     var tokenType = wrappedToken.Root.Descendants("{http://schemas.xmlsoap.org/ws/2005/02/trust}TokenType").First().Value; The decoded token contains the claims as key/value pairs, along with the issuer, audience (ACS realm), expiry date and an HMAC hash, which are in query string format. Separate them on the ampersand, and you can write out the claim values in your logged-in page:     var decoded = HttpUtility.UrlDecode(token);     foreach (var part in decoded.Split('&'))     {         Response.Write("<pre>" + part + "</pre><br/>");     } - which will produce something like this: http://schemas.microsoft.com/ws/2008/06/identity/claims/authenticationinstant=2012-08-31T06:57:01.855Z http://schemas.microsoft.com/ws/2008/06/identity/claims/authenticationmethod=http://schemas.microsoft.com/ws/2008/06/identity/authenticationmethod/windows http://schemas.microsoft.com/ws/2008/06/identity/claims/windowsaccountname=XYZ http://schemas.xmlsoap.org/ws/2005/05/identity/claims/[email protected] http://schemas.xmlsoap.org/ws/2005/05/identity/claims/[email protected] http://schemas.microsoft.com/accesscontrolservice/2010/07/claims/identityprovider=http://fs.svc.xyz.com/adfs/services/trust Audience=http://localhost/x.y.z ExpiresOn=1346402225 Issuer=https://x-y-z.accesscontrol.windows.net/ HMACSHA256=oDCeEDDAWEC8x+yBnTaCLnzp4L6jI0Z/xNK95PdZTts= The HMAC hash lets you validate the token to ensure it hasn’t been tampered with. You'll need the token signing key from ACS, then you can re-sign the token and compare hashes. There's a full implementation of an SWT parser and validator here: How To Request SWT Token From ACS And How To Validate It At The REST WCF Service Hosted In Windows Azure, and a cut-down claim inspector on my github code gallery: ACS Claim Inspector. Interestingly, ACS lets you have a value for your logged-in page which has no relation to the realm for authentication, so you can put this code into a generic claim inspector page, and set that to be your logged-in page for any relying party where you want to check what's being sent through. Particularly handy with ADFS, when you're modifying the claims provided, and want to quickly see the results.

    Read the article

  • Which of these algorithms is best for my goal?

    - by JonathonG
    I have created a program that restricts the mouse to a certain region based on a black/white bitmap. The program is 100% functional as-is, but uses an inaccurate, albeit fast, algorithm for repositioning the mouse when it strays outside the area. Currently, when the mouse moves outside the area, basically what happens is this: A line is drawn between a pre-defined static point inside the region and the mouse's new position. The point where that line intersects the edge of the allowed area is found. The mouse is moved to that point. This works, but only works perfectly for a perfect circle with the pre-defined point set in the exact center. Unfortunately, this will never be the case. The application will be used with a variety of rectangles and irregular, amorphous shapes. On such shapes, the point where the line drawn intersects the edge will usually not be the closest point on the shape to the mouse. I need to create a new algorithm that finds the closest point to the mouse's new position on the edge of the allowed area. I have several ideas about this, but I am not sure of their validity, in that they may have far too much overhead. While I am not asking for code, it might help to know that I am using Objective C / Cocoa, developing for OS X, as I feel the language being used might affect the efficiency of potential methods. My ideas are: Using a bit of trigonometry to project lines would work, but that would require some kind of intense algorithm to test every point on every line until it found the edge of the region... That seems too resource intensive since there could be something like 200 lines that would have each have to have as many as 200 pixels checked for black/white.... Using something like an A* pathing algorithm to find the shortest path to a black pixel; however, A* seems resource intensive, even though I could probably restrict it to only checking roughly in one direction. It also seems like it will take more time and effort than I have available to spend on this small portion of the much larger project I am working on, correct me if I am wrong and it would not be a significant amount of code (100 lines or around there). Mapping the border of the region before the application begins running the event tap loop. I think I could accomplish this by using my current line-based algorithm to find an edge point and then initiating an algorithm that checks all 8 pixels around that pixel, finds the next border pixel in one direction, and continues to do this until it comes back to the starting pixel. I could then store that data in an array to be used for the entire duration of the program, and have the mouse re-positioning method check the array for the closest pixel on the border to the mouse target position. That last method would presumably execute it's initial border mapping fairly quickly. (It would only have to map between 2,000 and 8,000 pixels, which means 8,000 to 64,000 checked, and I could even permanently store the data to make launching faster.) However, I am uncertain as to how much overhead it would take to scan through that array for the shortest distance for every single mouse move event... I suppose there could be a shortcut to restrict the number of elements in the array that will be checked to a variable number starting with the intersecting point on the line (from my original algorithm), and raise/lower that number to experiment with the overhead/accuracy tradeoff. Please let me know if I am over thinking this and there is an easier way that will work just fine, or which of these methods would be able to execute something like 30 times per second to keep mouse movement smooth, or if you have a better/faster method. I've posted relevant parts of my code below for reference, and included an example of what the area might look like. (I check for color value against a loaded bitmap that is black/white.) // // This part of my code runs every single time the mouse moves. // CGPoint point = CGEventGetLocation(event); float tX = point.x; float tY = point.y; if( is_in_area(tX,tY, mouse_mask)){ // target is inside O.K. area, do nothing }else{ CGPoint target; //point inside restricted region: float iX = 600; // inside x float iY = 500; // inside y // delta to midpoint between iX,iY and tX,tY float dX; float dY; float accuracy = .5; //accuracy to loop until reached do { dX = (tX-iX)/2; dY = (tY-iY)/2; if(is_in_area((tX-dX),(tY-dY),mouse_mask)){ iX += dX; iY += dY; } else { tX -= dX; tY -= dY; } } while (abs(dX)>accuracy || abs(dY)>accuracy); target = CGPointMake(roundf(tX), roundf(tY)); CGDisplayMoveCursorToPoint(CGMainDisplayID(),target); } Here is "is_in_area(int x, int y)" : bool is_in_area(NSInteger x, NSInteger y, NSBitmapImageRep *mouse_mask){ NSAutoreleasePool * pool = [[NSAutoreleasePool alloc] init]; NSUInteger pixel[4]; [mouse_mask getPixel:pixel atX:x y:y]; if(pixel[0]!= 0){ [pool release]; return false; } [pool release]; return true; }

    Read the article

  • Training a 'replacement', how to enforce standards?

    - by Mohgeroth
    Not sure that this is the right stack exchange site to ask this of, but here goes... Scope I work for a small company that employs a few hundred people. The development team for the company is small and works out of visual foxpro. A specific department in the company hired me in as a 'lone gunman' to fix and enhance a pre-existing invoicing system. I've successfully taken an Access application that suffered from a lot of risks and limitations and converted it into a C# application driven off of a SQL server backend. I have recently obtained my undergraduate and am no expert by any means. To help make up for that I've felt that earning microsoft certifications will force me to understand more about .net and how it functions. So, after giving my notice with 9 months in advance, 3 months ago a replacement finally showed up. Their role is to learn what I have been designing to an attempt to support the applications designed in C#. The Replacement Fresh out of college with no real-world work experience, the first instinct for anything involving data was and still is listboxes... any time data is mentioned the list box is the control of choice for the replacement. This has gotten to the point, no matter how many times I discuss other controls, where I've seen 5 listboxes on a single form. Classroom experience was almost all C++ console development. So, an example of where I have concern is in a winforms application: Users need to key Reasons into a table to select from later. Given that I know that a strongly typed data set exists, I can just drag the data source from the toolbox and it would create all of this for me. I realize this is a simple example but using databinding is the key. For the past few months now we have been talking about the strongly typed dataset, how to use it and where it interacts with other controls. Data sets, how they work in relation to binding sources, adapters and data grid views. After handing this project off I expected questions about how to implement these since for me this is the way to do it. What happened next simply floors me: An instance of an adapter from the strongly typed dataset was created in the activate event of the form, a table was created and filled with data. Then, a loop was made to manually add rows to a listbox from this table. Finally, a variable was kept to do lookups to figure out what ID the record was for updates if required. How do they modify records you ask? That was my first question too. You won't believe how simple it is, all you do it double click and they type into a pop-up prompt the new value to change it to. As a data entry operator, all the modal popups would drive me absolutely insane. The final solution exceeds 100 lines of code that must be maintained. So my concern is that none of this is sinking in... the department is only allowed 20 hours a week of their time. Up until last week, we've only been given 4-5 hours a week if I'm lucky. The past week or so, I've been lucky to get 10. Question WHAT DO I DO?! I have 4 weeks left until I leave and they fully 'support' this application. I love this job and the opportunity it has given me but it's time for me to spread my wings and find something new. I am in no way, shape or form convinced that they are ready to take over. I do feel that the replacement has the technical ability to 'figure it out' but instead of learning they just write code to do all of this stuff manually. If the replacement wants to code differently in the end, as long as it works I'm fine with that as horrifiying at it looks. However to support what I have designed they MUST to understand how it works and how I have used controls and the framework to make 'magic' happen. This project has about 40 forms, a database with over 30 some odd tables, triggers and stored procedures. It relates labor to invoices to contracts to projections... it's not as simple as it was three years ago when I began this project and the department is now in a position where they cannot survive without it. How in the world can I accomplish any of the following?: Enforce standards or understanding in constent design when the department manager keeps telling them they can do it however they want to Find a way to engage the replacement in active learning of the framework and system design that support must be given for Gracefully inform sr. management that 5-9 hours a week is simply not enough time to learn about the department, pre-existing processes, applications that need to be supported AND determine where potential enhancements to the system go... Yes I know this is a wall of text, thanks for reading through me but I simply don't know what I should be doing. For me, this job is a monster of a reference and things would look extremely bad if I left and things fell apart. How do I handle this?

    Read the article

  • When is my View too smart?

    - by Kyle Burns
    In this posting, I will discuss the motivation behind keeping View code as thin as possible when using patterns such as MVC, MVVM, and MVP.  Once the motivation is identified, I will examine some ways to determine whether a View contains logic that belongs in another part of the application.  While the concepts that I will discuss are applicable to most any pattern which favors a thin View, any concrete examples that I present will center on ASP.NET MVC. Design patterns that include a Model, a View, and other components such as a Controller, ViewModel, or Presenter are not new to application development.  These patterns have, in fact, been around since the early days of building applications with graphical interfaces.  The reason that these patterns emerged is simple – the code running closest to the user tends to be littered with logic and library calls that center around implementation details of showing and manipulating user interface widgets and when this type of code is interspersed with application domain logic it becomes difficult to understand and much more difficult to adequately test.  By removing domain logic from the View, we ensure that the View has a single responsibility of drawing the screen which, in turn, makes our application easier to understand and maintain. I was recently asked to take a look at an ASP.NET MVC View because the developer reviewing it thought that it possibly had too much going on in the view.  I looked at the .CSHTML file and the first thing that occurred to me was that it began with 40 lines of code declaring member variables and performing the necessary calculations to populate these variables, which were later either output directly to the page or used to control some conditional rendering action (such as adding a class name to an HTML element or not rendering another element at all).  This exhibited both of what I consider the primary heuristics (or code smells) indicating that the View is too smart: Member variables – in general, variables in View code are an indication that the Model to which the View is being bound is not sufficient for the needs of the View and that the View has had to augment that Model.  Notable exceptions to this guideline include variables used to hold information specifically related to rendering (such as a dynamically determined CSS class name or the depth within a recursive structure for indentation purposes) and variables which are used to facilitate looping through collections while binding. Arithmetic – as with member variables, the presence of arithmetic operators within View code are an indication that the Model servicing the View is insufficient for its needs.  For example, if the Model represents a line item in a sales order, it might seem perfectly natural to “normalize” the Model by storing the quantity and unit price in the Model and multiply these within the View to show the line total.  While this does seem natural, it introduces a business rule to the View code and makes it impossible to test that the rounding of the result meets the requirement of the business without executing the View.  Within View code, arithmetic should only be used for activities such as incrementing loop counters and calculating element widths. In addition to the two characteristics of a “Smart View” that I’ve discussed already, this View also exhibited another heuristic that commonly indicates to me the need to refactor a View and make it a bit less smart.  That characteristic is the existence of Boolean logic that either does not work directly with properties of the Model or works with too many properties of the Model.  Consider the following code and consider how logic that does not work directly with properties of the Model is just another form of the “member variable” heuristic covered earlier: @if(DateTime.Now.Hour < 12) {     <div>Good Morning!</div> } else {     <div>Greetings</div> } This code performs business logic to determine whether it is morning.  A possible refactoring would be to add an IsMorning property to the Model, but in this particular case there is enough similarity between the branches that the entire branching structure could be collapsed by adding a Greeting property to the Model and using it similarly to the following: <div>@Model.Greeting</div> Now let’s look at some complex logic around multiple Model properties: @if (ModelPageNumber + Model.NumbersToDisplay == Model.PageCount         || (Model.PageCount != Model.CurrentPage             && !Model.DisplayValues.Contains(Model.PageCount))) {     <div>There's more to see!</div> } In this scenario, not only is the View code difficult to read (you shouldn’t have to play “human compiler” to determine the purpose of the code), but it also complex enough to be at risk for logical errors that cannot be detected without executing the View.  Conditional logic that requires more than a single logical operator should be looked at more closely to determine whether the condition should be evaluated elsewhere and exposed as a single property of the Model.  Moving the logic above outside of the View and exposing a new Model property would simplify the View code to: @if(Model.HasMoreToSee) {     <div>There’s more to see!</div> } In this posting I have briefly discussed some of the more prominent heuristics that indicate a need to push code from the View into other pieces of the application.  You should now be able to recognize these symptoms when building or maintaining Views (or the Models that support them) in your applications.

    Read the article

  • [XSL-FO] Characters from other than English languages

    - by Lukasz Kurylo
    My client have departments in Europe Central and East, so there is highly possibility that in the generated pdfs there will be at least in the people names and/or surnames some specific characters for the country language.   With the XSL-FO we can use some out-of-the box fonts, e.g. the default is Times. We can change it for specific block of text or the entire document to other like Helvetica or Arial. All will be good to the moment that we use only an english alphabet. If we want to add e.g. some characters from polish or bulgarian language, in the *.fo file:         <fo:block >                 <fo:inline font-weight="bold">english: </fo:inline>                 <fo:inline font-weight="bold">yellow</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">polish: </fo:inline>                 <fo:inline font-weight="bold">zólty</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">russian: </fo:inline>                 <fo:inline font-weight="bold">??????</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">bulgarian: </fo:inline>                 <fo:inline font-weight="bold">????</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">english: </fo:inline>                 <fo:inline font-weight="bold">yellow</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">polish: </fo:inline>                 <fo:inline font-weight="bold"  font-family="Arial">zólty</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">russian: </fo:inline>                 <fo:inline font-weight="bold" font-family="Arial">??????</fo:inline>       </fo:block>       <fo:block>                 <fo:inline font-weight="bold">bulgarian: </fo:inline>                 <fo:inline font-weight="bold" font-family="Arial">????</fo:inline>       </fo:block>   The result can be diffrent from the expected depending on the selected font, e.g:                 As you can see Timer nor Arial work in this case.   The problem here is not related to XSL-FO, but rather to the renderer we are using. I have lost a lot of time to find a solution for the using by me XSL-FO –> PDF rendered to acquire these characters in my generated files. Fortunatelly all what have to be done it is to embed the font (or part of it) in the file(s) during rendering.   The renderer that I’m using it is an open source FO.NET.   For this one, the code to generate a pdf file looks that:   var fonet =  Fonet.FonetDriver.Make(); fonet.Render("source.fo", "result.pdf");   To emded the font in the pdf, we need to set the appropriate option to the driver:   fonet.Options = new Fonet.Render.Pdf.PdfRendererOptions() {       FontType = Fonet.Render.Pdf.FontType.Embed }; Right now, the pdf we get should look like this:               As you can see, the result for the Arial font looks exactly how it should, because this font has a characters included not only for the english language like the default Times, which we shouls avoid if we not generating a english-only documents.   This is worth to notice that in this situation the generated pdf file is quite large, it has more than 400 kb in size. This is of course because of embedding the entire font in it to make the document portable to systems, where the used font is not present. Instead on embedding the entire font, we can only embed the subset of used characters by changing the options to:   fonet.Options = new Fonet.Render.Pdf.PdfRendererOptions() {       FontType = Fonet.Render.Pdf.FontType.Subset };   Right now, this specific pdf is only 12 kb in size.

    Read the article

  • MVVM - how to make creating viewmodels at runtime less painfull

    - by Mr Happy
    I apologize for the long question, it reads a bit as a rant, but I promise it's not! I've summarized my question(s) below In the MVC world, things are straightforward. The Model has state, the View shows the Model, and the Controller does stuff to/with the Model (basically), a controller has no state. To do stuff the Controller has some dependencies on web services, repository, the lot. When you instantiate a controller you care about supplying those dependencies, nothing else. When you execute an action (method on Controller), you use those dependencies to retrieve or update the Model or calling some other domain service. If there's any context, say like some user wants to see the details of a particular item, you pass the Id of that item as parameter to the Action. Nowhere in the Controller is there any reference to any state. So far so good. Enter MVVM. I love WPF, I love data binding. I love frameworks that make data binding to ViewModels even easier (using Caliburn Micro a.t.m.). I feel things are less straightforward in this world though. Let's do the exercise again: the Model has state, the View shows the ViewModel, and the ViewModel does stuff to/with the Model (basically), a ViewModel does have state! (to clarify; maybe it delegates all the properties to one or more Models, but that means it must have a reference to the model one way or another, which is state in itself) To do stuff the ViewModel has some dependencies on web services, repository, the lot. When you instantiate a ViewModel you care about supplying those dependencies, but also the state. And this, ladies and gentlemen, annoys me to no end. Whenever you need to instantiate a ProductDetailsViewModel from the ProductSearchViewModel (from which you called the ProductSearchWebService which in turn returned IEnumerable<ProductDTO>, everybody still with me?), you can do one of these things: call new ProductDetailsViewModel(productDTO, _shoppingCartWebService /* dependcy */);, this is bad, imagine 3 more dependencies, this means the ProductSearchViewModel needs to take on those dependencies as well. Also changing the constructor is painfull. call _myInjectedProductDetailsViewModelFactory.Create().Initialize(productDTO);, the factory is just a Func, they are easily generated by most IoC frameworks. I think this is bad because Init methods are a leaky abstraction. You also can't use the readonly keyword for fields that are set in the Init method. I'm sure there are a few more reasons. call _myInjectedProductDetailsViewModelAbstractFactory.Create(productDTO); So... this is the pattern (abstract factory) that is usually recommended for this type of problem. I though it was genious since it satisfies my craving for static typing, until I actually started using it. The amount of boilerplate code is I think too much (you know, apart from the ridiculous variable names I get use). For each ViewModel that needs runtime parameters you'll get two extra files (factory interface and implementation), and you need to type the non-runtime dependencies like 4 extra times. And each time the dependencies change, you get to change it in the factory as well. It feels like I don't even use an DI container anymore. (I think Castle Windsor has some kind of solution for this [with it's own drawbacks, correct me if I'm wrong]). do something with anonymous types or dictionary. I like my static typing. So, yeah. Mixing state and behavior in this way creates a problem which don't exist at all in MVC. And I feel like there currently isn't a really adequate solution for this problem. Now I'd like to observe some things: People actually use MVVM. So they either don't care about all of the above, or they have some brilliant other solution. I haven't found an indepth example of MVVM with WPF. For example, the NDDD-sample project immensely helped me understand some DDD concepts. I'd really like it if someone could point me in the direction of something similar for MVVM/WPF. Maybe I'm doing MVVM all wrong and I should turn my design upside down. Maybe I shouldn't have this problem at all. Well I know other people have asked the same question so I think I'm not the only one. To summarize Am I correct to conclude that having the ViewModel being an integration point for both state and behavior is the reason for some difficulties with the MVVM pattern as a whole? Is using the abstract factory pattern the only/best way to instantiate a ViewModel in a statically typed way? Is there something like an in depth reference implementation available? Is having a lot of ViewModels with both state/behavior a design smell?

    Read the article

  • Windows for IoT, continued

    - by Valter Minute
    Originally posted on: http://geekswithblogs.net/WindowsEmbeddedCookbook/archive/2014/08/05/windows-for-iot-continued.aspxI received many interesting feedbacks on my previous blog post and I tried to find some time to do some additional tests. Bert Kleinschmidt pointed out that pins 2,3 and 10 of the Galileo are connected directly to the SOC, while pin 13, the one used for the sample sketch is controlled via an I2C I/O expander. I changed my code to use pin 2 instead of 13 (just changing the variable assignment at the beginning of the code) and latency was greatly reduced. Now each pulse lasts for 1.44ms, 44% more than the expected time, but ways better that the result we got using pin 13. I also used SetThreadPriority to increase the priority of the thread that was running the sketch to THREAD_PRIORITY_HIGHEST but that didn't change the results. When I was using the I2C-controlled pin I tried the same and the timings got ways worse (increasing more than 10 times) and so I did not commented on that part, wanting to investigate the issua a bit more in detail. It seems that increasing the priority of the application thread impacts negatively the I2C communication. I tried to use also the Linux-based implementation (using a different Galileo board since the one provided by MS seems to use a different firmware) and the results of running the sample blink sketch modified to use pin 2 and blink the led for 1ms are similar to those we got on the same board running Windows. Here the difference between expected time and measured time is worse, getting around 3.2ms instead of 1 (320% compared to 150% using Windows but far from the 100.1% we got with the 8-bit Arduino). Both systems were not under load during the test, maybe loading some applications that use part of the CPU time would make those timings even less reliable, but I think that those numbers are enough to draw some conclusions. It may not be worth running a full OS if what you need is Arduino compatibility. The Arduino UNO is probably the best Arduino you can find to perform this kind of development. The Galileo running the Linux-based stack or running Windows for IoT is targeted to be a platform for "Internet of Things" devices, whatever that means. At the moment I don't see the "I" part of IoT. We have low level interfaces (SPI, I2C, the GPIO pins) that can be used to connect sensors but the support for connectivity is limited and the amount of work required to deliver some data to the cloud (using a secure HTTP request or a message queuing system like APMQS or MQTT) is still big and the rich OS underneath seems to not provide any help doing that.Why should I use sockets and can't access all the high level connectivity features we have on "full" Windows?I know that it's possible to use some third party libraries, try to build them using the Windows For IoT SDK etc. but this means re-inventing the wheel every time and can also lead to some IP concerns if used for products meant to be closed-source. I hope that MS and Intel (and others) will focus less on the "coolness" of running (some) Arduino sketches and more on providing a better platform to people that really want to design devices that leverage internet connectivity and the cloud processing power to deliver better products and services. Providing a reliable set of connectivity services would be a great start. Providing support for .NET would be even better, leaving native code available for hardware access etc. I know that those components may require additional storage and memory etc. So making the OS componentizable (or, at least, provide a way to install additional components) would be a great way to let developers pick the parts of the system they need to develop their solution, knowing that they will integrate well together. I can understand that the Arduino and Raspberry Pi* success may have attracted the attention of marketing departments worldwide and almost any new development board those days is promoted as "XXX response to Arduino" or "YYYY alternative to Raspberry Pi", but this is misleading and prevents companies from focusing on how to deliver good products and how to integrate "IoT" features with their existing offer to provide, at the end, a better product or service to their customers. Marketing is important, but can't decide the key features of a product (the OS) that is going to be used to develop full products for end customers integrating it with hardware and application software. I really like the "hackable" nature of open-source devices and like to see that companies are getting more and more open in releasing information, providing "hackable" devices and supporting developers with documentation, good samples etc. On the other side being able to run a sketch designed for an 8 bit microcontroller on a full-featured application processor may sound cool and an easy upgrade path for people that just experimented with sensors etc. on Arduino but it's not, in my humble opinion, the main path to follow for people who want to deliver real products.   *Shameless self-promotion: if you are looking for a good book in Italian about the Raspberry Pi , try mine: http://www.amazon.it/Raspberry-Pi-alluso-Digital-LifeStyle-ebook/dp/B00GYY3OKO

    Read the article

  • How-to populate different select list content per table row

    - by frank.nimphius
    Normal 0 false false false EN-US X-NONE X-NONE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} A frequent requirement posted on the OTN forum is to render cells of a table column using instances of af:selectOneChoices with each af:selectOneChoice instance showing different list values. To implement this use case, the select list of the table column is populated dynamically from a managed bean for each row. The table's current rendered row object is accessible in the managed bean using the #{row} expression, where "row" is the value added to the table's var property. <af:table var="row">   ...   <af:column ...>     <af:selectOneChoice ...>         <f:selectItems value="#{browseBean.items}"/>     </af:selectOneChoice>   </af:column </af:table> The browseBean managed bean referenced in the code snippet above has a setItems and getItems method defined that is accessible from EL using the #{browseBean.items} expression. When the table renders, then the var property variable - the #{row} reference - is filled with the data object displayed in the current rendered table row. The managed bean getItems method returns a List<SelectItem>, which is the model format expected by the f:selectItems tag to populate the af:selectOneChoice list. public void setItems(ArrayList<SelectItem> items) {} //this method is executed for each table row public ArrayList<SelectItem> getItems() {   FacesContext fctx = FacesContext.getCurrentInstance();   ELContext elctx = fctx.getELContext();   ExpressionFactory efactory =          fctx.getApplication().getExpressionFactory();          ValueExpression ve =          efactory.createValueExpression(elctx, "#{row}", Object.class);      Row rw = (Row) ve.getValue(elctx);         //use one of the row attributes to determine which list to query and   //show in the current af:selectOneChoice list  // ...  ArrayList<SelectItem> alsi = new ArrayList<SelectItem>();  for( ... ){      SelectItem item = new SelectItem();        item.setLabel(...);        item.setValue(...);        alsi.add(item);   }   return alsi;} For better performance, the ADF Faces table stamps it data rows. Stamping means that the cell renderer component - af:selectOneChoice in this example - is instantiated once for the column and then repeatedly used to display the cell data for individual table rows. This however means that you cannot refresh a single select one choice component in a table to change its list values. Instead the whole table needs to be refreshed, rerunning the managed bean list query. Be aware that having individual list values per table row is an expensive operation that should be used only on small tables for Business Services with low latency data fetching (e.g. ADF Business Components and EJB) and with server side caching strategies for the queried data (e.g. storing queried list data in a managed bean in session scope).

    Read the article

  • Myths about Coding Craftsmanship part 2

    - by tom
    Myth 3: The source of all bad code is inept developers and stupid people When you review code is this what you assume?  Shame on you.  You are probably making assumptions in your code if you are assuming so much already.  Bad code can be the result of any number of causes including but not limited to using dated techniques (like boxing when generics are available), not following standards (“look how he does the spacing between arguments!” or “did he really just name that variable ‘bln_Hello_Cats’?”), being redundant, using properties, methods, or objects in a novel way (like switching on button.Text between “Hello World” and “Hello World “ //clever use of space character… sigh), not following the SOLID principals, hacking around assumptions made in earlier iterations / hacking in features that should be worked into the overall design.  The first two issues, while annoying are pretty easy to spot and can be fixed so easily.  If your coding team is made up of experienced professionals who are passionate about staying current then these shouldn’t be happening.  If you work with a variety of skills, backgrounds, and experience then there will be some of this stuff going on.  If you have an opportunity to mentor such a developer who is receptive to constructive criticism don’t be a jerk; help them and the codebase will improve.  A little patience can improve the codebase, your work environment, and even your perspective. The novelty and redundancy I have encountered has often been the use of creativity when language knowledge was perceived as unavailable or too time consuming.  When developers learn on the job you get a lot of this.  Rather than going to MSDN developers will use what they know.  Depending on the constraints of their assignment hacking together what they know may seem quite practical.  This was not stupid though I often wonder how much time is actually “saved” by hacking.  These issues are often harder to untangle if we ever do.  They can also grow out of control as we write hack after hack to make it work and get back to some development that is satisfying. Hacking upon an existing hack is what I call “feeding the monster”.  Code monsters are anti-patterns and hacks gone wild.  The reason code monsters continue to get bigger is that they keep growing in scope, touching more and more of the application.  This is not the result of dumb developers. It is probably the result of avoiding design, not taking the time to understand the problems or anticipate or communicate the vision of the product.  If our developers don’t understand the purpose of a feature or product how do we expect potential customers to do so? Forethought and organization are often what is missing from bad code.  Developers who do not use the SOLID principals should be encouraged to learn these principals and be given guidance on how to apply them.  The time “saved” by giving hackers room to hack will be made up for and then some. Not as technical debt but as shoddy work that if not replaced will be struggled with again and again.  Bad code is not the result of dumb developers (usually) it is the result of trying to do too much without the proper resources and neglecting the right thing that needs doing with the first thoughtless thing that comes into our heads. Object oriented code is all about relationships between objects.  Coders who believe their coworkers are all fools tend to write objects that are difficult to work with, not eager to explain themselves, and perform erratically and irrationally.  If you constantly find you are surrounded by idiots you may want to ask yourself if you are being unreasonable, if you are being closed minded, of if you have chosen the right profession.  Opening your mind up to the idea that you probably work with rational, well-intentioned people will probably make you a better coder and it might even make you less grumpy.  If you are surrounded by jerks who do not engage in the exchange of ideas who do not care about their customers or the durability of the code you are building together then I suggest you find a new place to work.  Myth 4: Customers don’t care about “beautiful” code Craftsmanship is customer focused because it means that the job was done right, the product will withstand the abuse, modifications, and scrutiny of our customers.  Users can appreciate a predictable timeline for a release, a product delivered on time and on budget, a feature set that does not interfere with the task(s) it is supporting, quick turnarounds on exception messages, self healing issues, and less issues.  These are all hindered by skimping on craftsmanship.  When we write data access and when we write reusable code.   What do you think?  Does bad code come primarily from low IQ individuals?  Do customers care about beautiful code?

    Read the article

  • boot issues - long delay, then "gave up waiting for root device"

    - by chazomaticus
    I've had this issue on and off for about two years now. I noticed it on a new (custom built) machine running 10.04 when that first came out, but then it went away until a few months ago. I've gone through a number of hard drive changes but I can't say specifically what if anything I changed hardware-wise to make it stop or start happening. I had assumed upgrading to a modern Ubuntu version would fix the issue, so I installed 12.04 beta on a spare partition last night, but it's still happening. Here's the issue. After grub loads and I select a kernel to boot, the screen goes blank save for a blinking cursor. It sits in this state for many long minutes before it finally gives up and gives me an initramfs shell with the message gave up waiting for root device (and lists the /dev/disk/by-uuid/... path it was waiting for) but no other specific diagnostic information. Now, here's the tricky part. For one, the problem is intermittent - sometimes it progresses from the blinking cursor to the Ubuntu splash boot screen in a few seconds, and once it gets that far it always continues booting fine. The really bizarre thing is that I can "force" it to "find" the root device by repeatedly pressing the space bar and hitting the machine's power button. If I tap those enough, eventually I will notice the hard drive light coming on, at which point it will always continue the boot process after a few seconds. Interestingly, if I wait slightly too long before pressing the power button (30s?), as soon as I press it I get the gave up waiting message and the initramfs shell. I've tried setting up /etc/fstab (and the grub menu.lst or whatever it's called nowadays) to use device names (e.g. /dev/sda1) instead of UUIDs, but I get the same effect just with the device name, not UUID, in the error message. I should also mention that when I boot to Windows 7, there is no issue. It boots slowly all the time just by virtue of being Windows, but it never hangs indefinitely. This would seem to indicate it's a problem in Ubuntu, not the hardware. It's pretty annoying to have to babysit the computer every time it boots. Any ideas? I'm at a loss. Not even sure how to diagnose the issue. Thanks! EDIT: Here's some dmesg output from 10.04. The 15 second gap is where it was doing nothing. I pressed the power button and space bar a few times, and the stuff at 16 seconds happened. Not sure what any of it means. [ 1.320250] scsi18 : ahci [ 1.320294] scsi19 : ahci [ 1.320320] ata19: SATA max UDMA/133 abar m8192@0xfd4fe000 port 0xfd4fe100 ir q 18 [ 1.320323] ata20: SATA max UDMA/133 abar m8192@0xfd4fe000 port 0xfd4fe180 ir q 18 [ 1.403886] usb 2-4: new high speed USB device using ehci_hcd and address 4 [ 1.562558] usb 2-4: configuration #1 chosen from 1 choice [ 16.477824] ata16: SATA link down (SStatus 0 SControl 300) [ 16.477843] ata19: SATA link down (SStatus 0 SControl 300) [ 16.477857] ata3: SATA link down (SStatus 0 SControl 300) [ 16.477895] ata15: SATA link down (SStatus 0 SControl 300) [ 16.477906] ata20: SATA link down (SStatus 0 SControl 300) [ 16.477977] ata17: SATA link down (SStatus 0 SControl 300) [ 16.478003] ata12: SATA link down (SStatus 0 SControl 300) [ 16.478046] ata13: SATA link down (SStatus 0 SControl 300) [ 16.478063] ata14: SATA link down (SStatus 0 SControl 300) [ 16.478108] ata11: SATA link down (SStatus 0 SControl 300) [ 16.478123] ata18: SATA link up 1.5 Gbps (SStatus 113 SControl 300) [ 16.478127] ata6: SATA link down (SStatus 0 SControl 300) [ 16.478157] ata5: SATA link down (SStatus 0 SControl 300) [ 16.478193] ata18.00: ATAPI: MARVELL VIRTUALL, 1.09, max UDMA/66 After that, it took its sweet time, and I had to keep hitting space bar to coax it along. Here's some more dmesg output from a little later in the boot process: [ 17.982291] input: BTC USB Multimedia Keyboard as /devices/pci0000:00/0000:00 :13.0/usb5/5-2/5-2:1.0/input/input4 [ 17.982335] generic-usb 0003:046E:5506.0002: input,hidraw1: USB HID v1.10 Key board [BTC USB Multimedia Keyboard] on usb-0000:00:13.0-2/input0 [ 18.005211] input: BTC USB Multimedia Keyboard as /devices/pci0000:00/0000:00 :13.0/usb5/5-2/5-2:1.1/input/input5 [ 18.005274] generic-usb 0003:046E:5506.0003: input,hiddev96,hidraw2: USB HID v1.10 Device [BTC USB Multimedia Keyboard] on usb-0000:00:13.0-2/input1 [ 22.484906] EXT4-fs (sda6): INFO: recovery required on readonly filesystem [ 22.484910] EXT4-fs (sda6): write access will be enabled during recovery [ 22.548542] EXT4-fs (sda6): recovery complete [ 22.549074] EXT4-fs (sda6): mounted filesystem with ordered data mode [ 32.516772] Adding 20482832k swap on /dev/sda5. Priority:-1 extents:1 across:20482832k [ 32.742540] udev: starting version 151 [ 33.002004] Bluetooth: Atheros AR30xx firmware driver ver 1.0 [ 33.008135] parport_pc 00:09: reported by Plug and Play ACPI [ 33.008186] parport0: PC-style at 0x378, irq 7 [PCSPP,TRISTATE] [ 33.012076] lp: driver loaded but no devices found [ 33.037271] ppdev: user-space parallel port driver [ 33.090256] lp0: using parport0 (interrupt-driven). Any clues in there?

    Read the article

  • From Pocket to Instapaper

    - by Michael Freidgeim
    Some time ago I’ve described the issues that I’ve had since a new version of Read It Later, named Pocket, was introduced.I’ve waited with hope for a new upgrade, but I had a huge disappointment with the latest version 16 June 2012. It didn’t fixed any of the two major problems, that I  experienced since new Pocket was introduced-  1. iPad app still didn’t show many of the saved links. 2. ability to rename articles on iPad still wasn’t restored.I’ve posted the message into their forum. They did not show my comment on their forum( I would name it censorship, not moderation), but a few days ago I’ve received an email, recommending “try logging out of the app on your iPad, and back in again.” Their suggestion helped,  but I don’t understand, why it is not posted as a recommendation on their support site.So I decided to try InstAPaper on my iPad, Previously I’ve used it for Kindle. I never considered it before on iPad, because there were no free demo and I was very satisfied with RIL free and then RIL Pro. Currently InstAPaper cost $3, so the price is not an issue.I’ve checked that it has most of features that I am using(e.g. renaming, folders) and I am quite happy with it now. Actually I am using Pocket (or RIL free) for old bookmarks( I have 1000+ stored on my iPad) and for new bookmarks I am using InstAPaper.Having a solid experience with RIL/Pocket I’ve created a list of suggestions to Marco Arment to implement.1. Some pages stored in InstAPaper have removed essential sections of the text. E.g in many blogs comments are not stored in  InstAPaper. Some pages lost almost all of important links (e.g. http://www.lib.rus.ec/a/32416 -sorry, in Russian). RIL/Pocket has 2 modes to store offline- Web view and Article view. Web View includes all links/images of the original page, but it’s very reliable. Article view suppose to strip unrelated information, but often corrupts the content. I prefer to use offline Web view.InstAPaper should also support offline Web view, in case if stripped view removes important part of content.2.  Black full screen Saving on iPad Safari is very annoying. After user pressed a bookmark, the saving has some delay and then for a few seconds prevents from reading the text.Would be better to show as message on the top part(as in Pocket ). I am surprised, that  a full screen popup was  implemented recently as a desired feature. 3.There are no comments allowed on http://blog.instapaper.com/. I would prefer to post some of these notes as comments on http://blog.instapaper.com/ rather than write them in my blog and then send link to Marco.(I found recommendation how to add support of comments on tumblr at http://www.tumblr.com/help, but then realized that Marko was the lead developer ofTumblr.)4. Also there is no support forum. I understand that maintenance of the forum ican be a hassle, but stackexchange fSome time ago I’ve described the issues that I’ve had since a new version of Read It Later, named Pocket, was introduced.I’ve waited with hope for a new upgrade, but I had a huge disappointment with the latest version 16 June 2012. It didn’t fixed any of the two major problems, that I  experienced since new Pocket was introduced- orums can be referred on  http://www.instapaper.com/main/support page, i.e.http://webapps.stackexchange.com/search?q=Instapaper  or http://apple.stackexchange.com/search?q=Instapaper 5. Tags are more convenient than folders. i.e. an ability for the same article to have more than one tag. Also creating of new folders is not supported offline, which is an annoying limitation.6. I would like to have a narrow list - additionally to existing list modes have a subject only list or subject+site list to show more list items on a screen.7. Limit of 500 offline articles sounds quite big, but my RIL list exceeded 1000, so it could be a issue in the future.8. Search button on iPad version is visible, but doesn’t work- it forces to buy Premium subscription. I think, that it’s not correct. If the button in a paid version is visible and enabled, it should  provide  a working functionality, e.g. search in article names only. And leave full-text search for the premium support.9..Copy URL is an important operation and deserves to be in a first level of Action menu, rather than in Share sub-menu.I’ve also have comment re post http://www.marco.org/2011/04/28/removed-instapaper-free. Marco Arment  explained, why he doesn’t provide free version of Instapaper.  I believe that he is loosing essential part of his customers. When I decided which of iPad application to choose, I’ve selected RIL, because I was able to play with free version, and I liked it. I didn’t have a chance to compare RIL and InstAPaper on iPad, so I’ve bought  RIL pro. For a user there is no point to pay even $3 , if there are similar free product, that user can try and see, is it suitable for him/her.I’ve also played with Readability. It doesn’t have folders or tags(which is very important for me), but nicely supports full text search

    Read the article

  • Migrating Core Data to new UIManagedDocument in iOS 5

    - by samerpaul
    I have an app that has been on the store since iOS 3.1, so there is a large install base out there that still uses Core Data loaded up in my AppDelegate. In the most recent set of updates, I raised the minimum version to 4.3 but still kept the same way of loading the data. Recently, I decided it's time to make the minimum version 5.1 (especially with 6 around the corner), so I wanted to start using the new fancy UIManagedDocument way of using Core Data. The issue with this though is that the old database file is still sitting in the iOS app, so there is no migrating to the new document. You have to basically subclass UIManagedDocument with a new model class, and override a couple of methods to do it for you. Here's a tutorial on what I did for my app TimeTag.  Step One: Add a new class file in Xcode and subclass "UIManagedDocument" Go ahead and also add a method to get the managedObjectModel out of this class. It should look like:   @interface TimeTagModel : UIManagedDocument   - (NSManagedObjectModel *)managedObjectModel;   @end   Step two: Writing the methods in the implementation file (.m) I first added a shortcut method for the applicationsDocumentDirectory, which returns the URL of the app directory.  - (NSURL *)applicationDocumentsDirectory {     return [[[NSFileManagerdefaultManager] URLsForDirectory:NSDocumentDirectoryinDomains:NSUserDomainMask] lastObject]; }   The next step was to pull the managedObjectModel file itself (.momd file). In my project, it's called "minimalTime". - (NSManagedObjectModel *)managedObjectModel {     NSString *path = [[NSBundlemainBundle] pathForResource:@"minimalTime"ofType:@"momd"];     NSURL *momURL = [NSURL fileURLWithPath:path];     NSManagedObjectModel *managedObjectModel = [[NSManagedObjectModel alloc] initWithContentsOfURL:momURL];          return managedObjectModel; }   After that, I need to check for a legacy installation and migrate it to the new UIManagedDocument file instead. This is the overridden method: - (BOOL)configurePersistentStoreCoordinatorForURL:(NSURL *)storeURL ofType:(NSString *)fileType modelConfiguration:(NSString *)configuration storeOptions:(NSDictionary *)storeOptions error:(NSError **)error {     // If legacy store exists, copy it to the new location     NSURL *legacyPersistentStoreURL = [[self applicationDocumentsDirectory] URLByAppendingPathComponent:@"minimalTime.sqlite"];          NSFileManager* fileManager = [NSFileManagerdefaultManager];     if ([fileManager fileExistsAtPath:legacyPersistentStoreURL.path])     {         NSLog(@"Old db exists");         NSError* thisError = nil;         [fileManager replaceItemAtURL:storeURL withItemAtURL:legacyPersistentStoreURL backupItemName:niloptions:NSFileManagerItemReplacementUsingNewMetadataOnlyresultingItemURL:nilerror:&thisError];     }          return [superconfigurePersistentStoreCoordinatorForURL:storeURL ofType:fileType modelConfiguration:configuration storeOptions:storeOptions error:error]; }   Basically what's happening above is that it checks for the minimalTime.sqlite file inside the app's bundle on the iOS device.  If the file exists, it tells you inside the console, and then tells the fileManager to replace the storeURL (inside the method parameter) with the legacy URL. This basically gives your app access to all the existing data the user has generated (otherwise they would load into a blank app, which would be disastrous). It returns a YES if successful (by calling it's [super] method). Final step: Actually load this database Due to how my app works, I actually have to load the database at launch (instead of shortly after, which would be ideal). I call a method called loadDatabase, which looks like this: -(void)loadDatabase {     static dispatch_once_t onceToken;          // Only do this once!     dispatch_once(&onceToken, ^{         // Get the URL         // The minimalTimeDB name is just something I call it         NSURL *url = [[selfapplicationDocumentsDirectory] URLByAppendingPathComponent:@"minimalTimeDB"];         // Init the TimeTagModel (our custom class we wrote above) with the URL         self.timeTagDB = [[TimeTagModel alloc] initWithFileURL:url];           // Setup the undo manager if it's nil         if (self.timeTagDB.undoManager == nil){             NSUndoManager *undoManager = [[NSUndoManager  alloc] init];             [self.timeTagDB setUndoManager:undoManager];         }                  // You have to actually check to see if it exists already (for some reason you can't just call "open it, and if it's not there, create it")         if ([[NSFileManagerdefaultManager] fileExistsAtPath:[url path]]) {             // If it does exist, try to open it, and if it doesn't open, let the user (or at least you) know!             [self.timeTagDB openWithCompletionHandler:^(BOOL success){                 if (!success) {                     // Handle the error.                     NSLog(@"Error opening up the database");                 }                 else{                     NSLog(@"Opened the file--it already existed");                     [self refreshData];                 }             }];         }         else {             // If it doesn't exist, you need to attempt to create it             [self.timeTagDBsaveToURL:url forSaveOperation:UIDocumentSaveForCreatingcompletionHandler:^(BOOL success){                 if (!success) {                     // Handle the error.                     NSLog(@"Error opening up the database");                 }                 else{                     NSLog(@"Created the file--it did not exist");                     [self refreshData];                 }             }];         }     }); }   If you're curious what refreshData looks like, it sends out a NSNotification that the database has been loaded: -(void)refreshData {     NSNotification* refreshNotification = [NSNotificationnotificationWithName:kNotificationCenterRefreshAllDatabaseData object:self.timeTagDB.managedObjectContext  userInfo:nil];     [[NSNotificationCenter defaultCenter] postNotification:refreshNotification];     }   The kNotificationCenterRefreshAllDatabaseData is just a constant I have defined elsewhere that keeps track of all the NSNotification names I use. I pass the managedObjectContext of the newly created file so that my view controllers can have access to it, and start passing it around to one another. The reason we do this as a Notification is because this is being run in the background, so we can't know exactly when it finishes. Make sure you design your app for this! Have some kind of loading indicator, or make sure your user can't attempt to create a record before the database actually exists, because it will crash the app.

    Read the article

< Previous Page | 645 646 647 648 649 650 651 652 653 654 655 656  | Next Page >