Search Results

Search found 5885 results on 236 pages for 'finally'.

Page 206/236 | < Previous Page | 202 203 204 205 206 207 208 209 210 211 212 213  | Next Page >

  • Mscorlib mocking minus the attribute

    - by mehfuzh
    Mocking .net framework members (a.k.a. mscorlib) is always a daunting task. It’s the breed of static and final methods and full of surprises. Technically intercepting mscorlib members is completely different from other class libraries. This is the reason it is dealt differently. Generally, I prefer writing a wrapper around an mscorlib member (Ex. File.Delete(“abc.txt”)) and expose it via interface but that is not always an easy task if you already have years old codebase. While mocking mscorlib members first thing that comes to people’s mind is DateTime.Now. If you Google through, you will find tons of example dealing with just that. May be it’s the most important class that we can’t ignore and I will create an example using JustMock Q2 with the same. In Q2 2012, we just get rid of the MockClassAtrribute for mocking mscorlib members. JustMock is already attribute free for mocking class libraries. We radically think that vendor specific attributes only makes your code smelly and therefore decided the same for mscorlib. Now, I want to fake DateTime.Now for the following class: public class NestedDateTime { public DateTime GetDateTime() { return DateTime.Now; } } It is the simplest one that can be. The first thing here is that I tell JustMock “hey we have a DateTime.Now in NestedDateTime class that we want to mock”. To do so, during the test initialization I write this: .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Mock.Replace(() => DateTime.Now).In<NestedDateTime>(x => x.GetDateTime());.csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } I can also define it for all the members in the class, but that’s just a waste of extra watts. Mock.Replace(() => DateTime.Now).In<NestedDateTime>(); Now question, why should I bother doing it? The answer is that I am not using attribute and with this approach, I can mock any framework members not just File, FileInfo or DateTime. Here to note that we already mock beyond the three but when nested around a complex class, JustMock was not intercepting it correctly. Therefore, we decided to get rid of the attribute altogether fixing the issue. Finally, I write my test as usual. [TestMethod] public void ShouldAssertMockingDateTimeFromNestedClass() { var expected = new DateTime(2000, 1, 1); Mock.Arrange(() => DateTime.Now).Returns(expected); Assert.Equal(new NestedDateTime().GetDateTime(), expected); } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } That’s it, we are good. Now let me do the same for a random one, let’s say I want mock a member from DriveInfo: Mock.Replace<DriveInfo[]>(() => DriveInfo.GetDrives()).In<MsCorlibFixture>(x => x.ShouldReturnExpectedDriveWhenMocked()); Moving forward, I write my test: [TestMethod] public void ShouldReturnExpectedDriveWhenMocked() { Mock.Arrange(() => DriveInfo.GetDrives()).MustBeCalled(); DriveInfo.GetDrives(); Mock.Assert(()=> DriveInfo.GetDrives()); } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Here is one convention; you have to replace the mscorlib member before executing the target method that contains it. Here the call to DriveInfo is within the MsCorlibFixture therefore it should be defined during test initialization or before executing the test method. Hope this gives you the idea.

    Read the article

  • SQL SERVER – Weekly Series – Memory Lane – #031

    - by Pinal Dave
    Here is the list of selected articles of SQLAuthority.com across all these years. Instead of just listing all the articles I have selected a few of my most favorite articles and have listed them here with additional notes below it. Let me know which one of the following is your favorite article from memory lane. 2007 Find Table without Clustered Index – Find Table with no Primary Key Clustered index is very important concept for any table. They impact the performance very heavily. Here is a quick script to find tables without a clustered index. Replace TEXT with VARCHAR(MAX) – Stop using TEXT, NTEXT, IMAGE Data Types Question: “Is VARCHAR (MAX) big enough to store the TEXT field?” Answer: “Yes, VARCHAR(MAX) is big enough to accommodate TEXT field. TEXT, NTEXT and IMAGE data types of SQL Server 2000 will be deprecated in a future version of SQL Server, SQL Server 2005 provides backward compatibility to data types but it is recommended to use new data types which are VARHCAR (MAX), NVARCHAR (MAX) and VARBINARY (MAX).” Limiting Result Sets by Using TABLESAMPLE – Examples Introduced in SQL Server 2005, TABLESAMPLE allows you to extract a sampling of rows from a table in the FROM clause. The rows retrieved are random and they are are not in any order. This sampling can be based on a percentage of number of rows. You can use TABLESAMPLE when only a sampling of rows is necessary for the application instead of a full result set. User Defined Functions (UDF) Limitations UDF have its own advantage and usage but in this article we will see the limitation of UDF. Things UDF can not do and why Stored Procedure are considered as more flexible then UDFs. Stored Procedure are more flexibility then User Defined Functions(UDF). However, this blog post is a good read to know what are the limitations of UDF. Change Database Compatible Level – Backward Compatibility For a long time SQL Server stayed on the compatibility level of 80 which is of SQL Server 2000. However, as soon as SQL Server 2005 introduced the issue of compatibility was quite a major issue. Since that time MS has been releasing the versions at every 2-3 years, changing compatibility is a ever popular topic. In this blog post, we learn how we can do the same using T-SQL. We can also do the same using SSMS and here is the blog post for the same: Change Database Compatible Level – Backward Compatibility – Part 2 – Management Studio. Constraint on VARCHAR(MAX) Field To Limit It Certain Length How can I limit the VARCHAR(MAX) field with maximum length of 12500 characters only. His Question was valid as our application was allowed 12500 characters. First of all – this requirement is bit strange but if someone wants to do the same, they can do it as described in this blog post. 2008 UNPIVOT Table Example Understanding UNPIVOT can be very complicated at times. In this blog post, I have attempted to explain the same concept in very simple words. Create Default Constraint Over Table Column A simple straight to script blog post – I still use this blog quite many times for my own reference. UDF – Get the Day of the Week Function It took me 4 iteration to find this very simple function which can immediately get the day of the week in a single line. 2009 Find Hostname and Current Logged In User Name There are two tricks listed in this blog post where users can find out the hostname and current logged user name immediately and very easily. Interesting Observation of Logon Trigger On All Servers When I was doing a project, I made an interesting observation of executing a logon trigger multiple times. It was absolutely unexpected for me! As I was logging only once, naturally, I was expecting the entry only once. However, it did it multiple times on different threads – indeed an eccentric phenomenon at first sight! Difference Between Candidate Keys and Primary Key One needs to be very careful in selecting the Primary Key as an incorrect selection can adversely impact the database architect and future normalization. For a Candidate Key to qualify as a Primary Key, it should be Non-NULL and unique in any domain. I have observed quite often that Primary Keys are seldom changed. I would like to have your feedback on not changing a Primary Key. Create Multiple Filegroup For Single Database Why should one create multiple file group for any database and what are the advantages of the same. In this blog post, I explain the same in detail. List All Objects Created on All Filegroups in Database In this blog post we discuss the essential question – “How can I find which object belongs to which filegroup. Is there any way to know this?” 2010 DATE and TIME in SQL Server 2008 When DATE is converted to DATETIME it adds the of midnight. When TIME is converted to DATETIME it adds the date of 1900 and it is something one wants to consider if you are going to run scripts from SQL Server 2008 to earlier version with CONVERT. Disabled Index and Update Statistics If you do not need a nonclustered index, I suggest you to drop it as keeping them disabled is an overhead on your system. This is because every time the statistics are updated for system all the statistics for disabled indexes are also updated. Precision of SMALLDATETIME – A 1 Minute Precision The precision of the datatype SMALLDATETIME is 1 minute. It discards the seconds by rounding up or rounding down any seconds greater than zero. 2011 Getting Columns Headers without Result Data – SET FMTONLY ON SET FMTONLY ON returns only metadata to the client. It can be used to test the format of the response without actually running the query. When this setting is ON the resultset only have headers of the results but no data. Copy Database from Instance to Another Instance – Copy Paste in SQL Server SQL Server has a feature which copy database from one database to another database and it can be automated as well using SSIS. Make sure you have SQL Server Agent Turned on as this feature will create a job. Puzzle – SELECT * vs SELECT COUNT(*) If you have ever wondered SELECT * gives error when executed alone but SELECT COUNT(*) does not. Why? in that case, you should read this blog post. Creating All New Database with Full Recovery Model This blog post is very based on very interesting story where the user wants to do something by default for every single new database created. Model database is a secret weapon which should be used very carefully and with proper evalution. If used carefully this can be a very much beneficiary when we need a newly created database behave in certain fashion. 2012 In year 2012 I had two interesting series ran on the blog. If there is no fun in learning, the learning becomes a burden. For the same reason, I had decided to build a three part quiz around SEQUENCE. The quiz was to identify the next value of the sequence. I encourage all of you to take part in this fun quiz. Guess the Next Value – Puzzle 1 Guess the Next Value – Puzzle 2 Guess the Next Value – Puzzle 3 Can anyone remember their final day of schooling?  This is probably a silly question because – of course you can!  Many people mark this as the most exciting, happiest day of their life.  It marks the end of testing, the end of following rules set by teachers, and the beginning of finally being able to earn money and work in your chosen field. Read five part series on developer training subject Developer Training - Importance and Significance - Part 1 Developer Training – Employee Morals and Ethics – Part 2 Developer Training – Difficult Questions and Alternative Perspective - Part 3 Developer Training – Various Options for Developer Training – Part 4 Developer Training – A Conclusive Summary- Part 5 Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Memory Lane, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

    Read the article

  • Why people don't patch and upgrade?!?

    - by Mike Dietrich
    Discussing the topic "Why Upgrade" or "Why not Upgrade" is not always fun. Actually the arguments repeat from customer to customer. Typically we hear things such as: A PSU or Patch Set introduces new bugs A new PSU or Patch Set introduces new features which lead to risk and require application verification  Patching means risk Patching changes the execution plans Patching requires too much testing Patching is too much work for our DBAs Patching costs a lot of money and doesn't pay out And to be very honest sometimes it's hard for me to stay calm in such discussions. Let's discuss some of these points a bit more in detail. A PSU or Patch Set introduces new bugsWell, yes, that is true as no software containing more than some lines of code is bug free. This applies to Oracle's code as well as too any application or operating system code. But first of all, does that mean you never patch your OS because the patch may introduce new flaws? And second, what is the point of saying "it introduces new bugs"? Does that mean you will never get rid of the mean issues we know about and we fixed already? Scroll down from MOS Note:161818.1 to the patch release you are on, no matter if it's 10.2.0.4 or 11.2.0.3 and check for the Known Issues And Alerts.Will you take responsibility to know about all these issues and refuse to upgrade to 11.2.0.4? I won't. A new PSU or Patch Set introduces new featuresOk, we can discuss that. Offering new functionality within a database patch set is a dubious thing. It has advantages such as in 11.2.0.4 where we backported Database Redaction to. But this is something you will only use once you have an Advanced Security license. I interpret that statement I've heard quite often from customers in a different way: People don't want to get surprises such as new behaviour. This certainly gives everybody a hard time. And we've had many examples in the past (SESSION_CACHED_CURSROS in 10.2.0.4,  _DATAFILE_WRITE_ERRORS_CRASH_INSTANCE in 11.2.0.2 and others) where those things weren't documented, not even in the README. Thanks to many friends out there I learned about those as well. So new behaviour is the topic people consider as risky - not really new features. And just to point this out: A PSU never brings in new features or new behaviour by definition! Patching means riskDoes it really mean risk? Yes, there were issues in the past (and sometimes in the present as well) where a patch didn't get installed correctly. But personally I consider it way more risky to not patch. Keep that in mind: The day Oracle publishes an PSU (or CPU) containing security fixes all the great security experts out there go public with their findings as well. So from that day on even my grandma can find out about those issues and try to attack somebody. Now a lot of people say: "My database does not face the internet." And I will answer: "The enemy is sitting already behind your firewalls. And knows potentially about these things." My statement: Not patching introduces way more risk to your environment than patching. Seriously! Patching changes the execution plansDo they really? I agree - there's a very small risk for this happening with Patch Sets. But not with PSUs or CPUs as they contain no optimizer fixes changing behaviour (but they may contain fixes curing wrong-query-result-bugs). But what's the point of a changing execution plan? In Oracle Database 11g it is so simple to be prepared. SQL Plan Management is a free EE feature - so once that occurs you'll put the plan into the Plan Baseline. Basta! Yes, you wouldn't like to get such surprises? Than please use the SQL Performance Analyzer (SPA) from Real Application Testing and you'll detect that easily upfront in minutes. And not to forget this, a plan change can also be very positive!Yes, there's a little risk with a database patchset - and we have many possibilites to detect this before patching. Patching requires too much testingWell, does it really? I have seen in the past 12 years how people test. There are very different efforts and approaches on this. I have seen people spending a hell of money on licenses or on project team staffing. And I have seen people sailing blindly without any tests just going the John-Wayne-approach.Proper tools will allow you to test easily without too much efforts. See the paragraph above. We have used Real Application Testing in so many customer projects reducing the amount of work spend on testing by over 50%. But apart from that at some point you will have to stop testing. If you don't you'll get lost and you'll burn money. There's no 100% guaranty. You will have to deal with a little risk as reaching the final 5% of certainty will cost you the same as it did cost to reach 95%. And doing this will lead to abnormal long product cycles that you'll run behind forever. And this will cost even more money. Patching is too much work for our DBAsPatching is a lot of work. I agree. And it's no fun work. It's boring, annoying. You don't learn much from that. That's why you should try to automate this task. Use the Database's Lifecycle Management Pack. And don't cry about the fact that it costs money. Yes it does. But it will ease the process and you'll save a lot of costs as you don't waste your valuable time with patching. Or use Oracle Database 12c Oracle Multitenant and patch either by unplug/plug or patch an entire container database with all PDBs with one patch in one task. We have customer reference cases proofing it saved them 75% of time, effort and cost since they've used Lifecycle Management Pack. So why don't you use it? Patching costs a lot of money and doesn't pay outWell, see my statements in the paragraph above. And it pays out as flying with a database with 100 known critical flaws in it which are already fixed by Oracle (such as in the Oct 2013 PSU for Oracle Database 12c) will cost ways more in case of failure or even data loss. Bet with me? Let me finally ask you some questions. What cell phone are you using and which OS does it run? Do you have an iPhone 5 and did you upgrade already to iOS 7.0.3? I've just encountered on mine that the alarm (which I rely on when traveling) has gotten now a dependency on the physical switch "sound on/off". If it is switched to "off" physically the alarm rings "silently". What a wonderful example of a behaviour change coming in with a patch set. Will this push you to stay with iOS5 or iOS6? No, because those have security flaws which won't be fixed anymore. What browser are you surfing with? Do you use Mozilla 3.6? Well, congratulations to all the hackers. It will be easy for them to attack you and harm your system. I'd guess you have the auto updater on.  Same for Google Chrome, Safari, IE. Right? -Mike The T.htmtableborders, .htmtableborders td, .htmtableborders th {border : 1px dashed lightgrey ! important;} html, body { border: 0px; } body { background-color: #ffffff; } img, hr { cursor: default }

    Read the article

  • General Purpose ASP.NET Data Source Control

    - by Ricardo Peres
    OK, you already know about the ObjectDataSource control, so what’s wrong with it? Well, for once, it doesn’t pass any context to the SelectMethod, you only get the parameters supplied on the SelectParameters plus the desired ordering, starting page and maximum number of rows to display. Also, you must have two separate methods, one for actually retrieving the data, and the other for getting the total number of records (SelectCountMethod). Finally, you don’t get a chance to alter the supplied data before you bind it to the target control. I wanted something simple to use, and more similar to ASP.NET 4.5, where you can have the select method on the page itself, so I came up with CustomDataSource. Here’s how to use it (I chose a GridView, but it works equally well with any regular data-bound control): 1: <web:CustomDataSourceControl runat="server" ID="datasource" PageSize="10" OnData="OnData" /> 2: <asp:GridView runat="server" ID="grid" DataSourceID="datasource" DataKeyNames="Id" PageSize="10" AllowPaging="true" AllowSorting="true" /> The OnData event handler receives a DataEventArgs instance, which contains some properties that describe the desired paging location and size, and it’s where you return the data plus the total record count. Here’s a quick example: 1: protected void OnData(object sender, DataEventArgs e) 2: { 3: //just return some data 4: var data = Enumerable.Range(e.StartRowIndex, e.PageSize).Select(x => new { Id = x, Value = x.ToString(), IsPair = ((x % 2) == 0) }); 5: e.Data = data; 6: //the total number of records 7: e.TotalRowCount = 100; 8: } Here’s the code for the DataEventArgs: 1: [Serializable] 2: public class DataEventArgs : EventArgs 3: { 4: public DataEventArgs(Int32 pageSize, Int32 startRowIndex, String sortExpression, IOrderedDictionary parameters) 5: { 6: this.PageSize = pageSize; 7: this.StartRowIndex = startRowIndex; 8: this.SortExpression = sortExpression; 9: this.Parameters = parameters; 10: } 11:  12: public IEnumerable Data 13: { 14: get; 15: set; 16: } 17:  18: public IOrderedDictionary Parameters 19: { 20: get; 21: private set; 22: } 23:  24: public String SortExpression 25: { 26: get; 27: private set; 28: } 29:  30: public Int32 StartRowIndex 31: { 32: get; 33: private set; 34: } 35:  36: public Int32 PageSize 37: { 38: get; 39: private set; 40: } 41:  42: public Int32 TotalRowCount 43: { 44: get; 45: set; 46: } 47: } As you can guess, the StartRowIndex and PageSize receive the starting row and the desired page size, where the page size comes from the PageSize property on the markup. There’s also a SortExpression, which gets passed the sorted-by column and direction (if descending) and a dictionary containing all the values coming from the SelectParameters collection, if any. All of these are read only, and it is your responsibility to fill in the Data and TotalRowCount. The code for the CustomDataSource is very simple: 1: [NonVisualControl] 2: public class CustomDataSourceControl : DataSourceControl 3: { 4: public CustomDataSourceControl() 5: { 6: this.SelectParameters = new ParameterCollection(); 7: } 8:  9: protected override DataSourceView GetView(String viewName) 10: { 11: return (new CustomDataSourceView(this, viewName)); 12: } 13:  14: internal void GetData(DataEventArgs args) 15: { 16: this.OnData(args); 17: } 18:  19: protected virtual void OnData(DataEventArgs args) 20: { 21: EventHandler<DataEventArgs> data = this.Data; 22:  23: if (data != null) 24: { 25: data(this, args); 26: } 27: } 28:  29: [Browsable(false)] 30: [DesignerSerializationVisibility(DesignerSerializationVisibility.Visible)] 31: [PersistenceMode(PersistenceMode.InnerProperty)] 32: public ParameterCollection SelectParameters 33: { 34: get; 35: private set; 36: } 37:  38: public event EventHandler<DataEventArgs> Data; 39:  40: public Int32 PageSize 41: { 42: get; 43: set; 44: } 45: } Also, the code for the accompanying internal – as there is no need to use it from outside of its declaring assembly - data source view: 1: sealed class CustomDataSourceView : DataSourceView 2: { 3: private readonly CustomDataSourceControl dataSourceControl = null; 4:  5: public CustomDataSourceView(CustomDataSourceControl dataSourceControl, String viewName) : base(dataSourceControl, viewName) 6: { 7: this.dataSourceControl = dataSourceControl; 8: } 9:  10: public override Boolean CanPage 11: { 12: get 13: { 14: return (true); 15: } 16: } 17:  18: public override Boolean CanRetrieveTotalRowCount 19: { 20: get 21: { 22: return (true); 23: } 24: } 25:  26: public override Boolean CanSort 27: { 28: get 29: { 30: return (true); 31: } 32: } 33:  34: protected override IEnumerable ExecuteSelect(DataSourceSelectArguments arguments) 35: { 36: IOrderedDictionary parameters = this.dataSourceControl.SelectParameters.GetValues(HttpContext.Current, this.dataSourceControl); 37: DataEventArgs args = new DataEventArgs(this.dataSourceControl.PageSize, arguments.StartRowIndex, arguments.SortExpression, parameters); 38:  39: this.dataSourceControl.GetData(args); 40:  41: arguments.TotalRowCount = args.TotalRowCount; 42: arguments.MaximumRows = this.dataSourceControl.PageSize; 43: arguments.AddSupportedCapabilities(DataSourceCapabilities.Page | DataSourceCapabilities.Sort | DataSourceCapabilities.RetrieveTotalRowCount); 44: arguments.RetrieveTotalRowCount = true; 45:  46: if (!(args.Data is ICollection)) 47: { 48: return (args.Data.OfType<Object>().ToList()); 49: } 50: else 51: { 52: return (args.Data); 53: } 54: } 55: } As always, looking forward to hearing from you!

    Read the article

  • Why do we (really) program to interfaces?

    - by Kyle Burns
    One of the earliest lessons I was taught in Enterprise development was "always program against an interface".  This was back in the VB6 days and I quickly learned that no code would be allowed to move to the QA server unless my business objects and data access objects each are defined as an interface and have a matching implementation class.  Why?  "It's more reusable" was one answer.  "It doesn't tie you to a specific implementation" a slightly more knowing answer.  And let's not forget the discussion ending "it's a standard".  The problem with these responses was that senior people didn't really understand the reason we were doing the things we were doing and because of that, we were entirely unable to realize the intent behind the practice - we simply used interfaces and had a bunch of extra code to maintain to show for it. It wasn't until a few years later that I finally heard the term "Inversion of Control".  Simply put, "Inversion of Control" takes the creation of objects that used to be within the control (and therefore a responsibility of) of your component and moves it to some outside force.  For example, consider the following code which follows the old "always program against an interface" rule in the manner of many corporate development shops: 1: ICatalog catalog = new Catalog(); 2: Category[] categories = catalog.GetCategories(); In this example, I met the requirement of the rule by declaring the variable as ICatalog, but I didn't hit "it doesn't tie you to a specific implementation" because I explicitly created an instance of the concrete Catalog object.  If I want to test the functionality of the code I just wrote I have to have an environment in which Catalog can be created along with any of the resources upon which it depends (e.g. configuration files, database connections, etc) in order to test my functionality.  That's a lot of setup work and one of the things that I think ultimately discourages real buy-in of unit testing in many development shops. So how do I test my code without needing Catalog to work?  A very primitive approach I've seen is to change the line the instantiates catalog to read: 1: ICatalog catalog = new FakeCatalog();   once the test is run and passes, the code is switched back to the real thing.  This obviously poses a huge risk for introducing test code into production and in my opinion is worse than just keeping the dependency and its associated setup work.  Another popular approach is to make use of Factory methods which use an object whose "job" is to know how to obtain a valid instance of the object.  Using this approach, the code may look something like this: 1: ICatalog catalog = CatalogFactory.GetCatalog();   The code inside the factory is responsible for deciding "what kind" of catalog is needed.  This is a far better approach than the previous one, but it does make projects grow considerably because now in addition to the interface, the real implementation, and the fake implementation(s) for testing you have added a minimum of one factory (or at least a factory method) for each of your interfaces.  Once again, developers say "that's too complicated and has me writing a bunch of useless code" and quietly slip back into just creating a new Catalog and chalking any test failures up to "it will probably work on the server". This is where software intended specifically to facilitate Inversion of Control comes into play.  There are many libraries that take on the Inversion of Control responsibilities in .Net and most of them have many pros and cons.  From this point forward I'll discuss concepts from the standpoint of the Unity framework produced by Microsoft's Patterns and Practices team.  I'm primarily focusing on this library because it questions about it inspired this posting. At Unity's core and that of most any IoC framework is a catalog or registry of components.  This registry can be configured either through code or using the application's configuration file and in the most simple terms says "interface X maps to concrete implementation Y".  It can get much more complicated, but I want to keep things at the "what does it do" level instead of "how does it do it".  The object that exposes most of the Unity functionality is the UnityContainer.  This object exposes methods to configure the catalog as well as the Resolve<T> method which is used to obtain an instance of the type represented by T.  When using the Resolve<T> method, Unity does not necessarily have to just "new up" the requested object, but also can track dependencies of that object and ensure that the entire dependency chain is satisfied. There are three basic ways that I have seen Unity used within projects.  Those are through classes directly using the Unity container, classes requiring injection of dependencies, and classes making use of the Service Locator pattern. The first usage of Unity is when classes are aware of the Unity container and directly call its Resolve method whenever they need the services advertised by an interface.  The up side of this approach is that IoC is utilized, but the down side is that every class has to be aware that Unity is being used and tied directly to that implementation. Many developers don't like the idea of as close a tie to specific IoC implementation as is represented by using Unity within all of your classes and for the most part I agree that this isn't a good idea.  As an alternative, classes can be designed for Dependency Injection.  Dependency Injection is where a force outside the class itself manipulates the object to provide implementations of the interfaces that the class needs to interact with the outside world.  This is typically done either through constructor injection where the object has a constructor that accepts an instance of each interface it requires or through property setters accepting the service providers.  When using dependency, I lean toward the use of constructor injection because I view the constructor as being a much better way to "discover" what is required for the instance to be ready for use.  During resolution, Unity looks for an injection constructor and will attempt to resolve instances of each interface required by the constructor, throwing an exception of unable to meet the advertised needs of the class.  The up side of this approach is that the needs of the class are very clearly advertised and the class is unaware of which IoC container (if any) is being used.  The down side of this approach is that you're required to maintain the objects passed to the constructor as instance variables throughout the life of your object and that objects which coordinate with many external services require a lot of additional constructor arguments (this gets ugly and may indicate a need for refactoring). The final way that I've seen and used Unity is to make use of the ServiceLocator pattern, of which the Patterns and Practices team has also provided a Unity-compatible implementation.  When using the ServiceLocator, your class calls ServiceLocator.Retrieve in places where it would have called Resolve on the Unity container.  Like using Unity directly, it does tie you directly to the ServiceLocator implementation and makes your code aware that dependency injection is taking place, but it does have the up side of giving you the freedom to swap out the underlying IoC container if necessary.  I'm not hugely concerned with hiding IoC entirely from the class (I view this as a "nice to have"), so the single biggest problem that I see with the ServiceLocator approach is that it provides no way to proactively advertise needs in the way that constructor injection does, allowing more opportunity for difficult to track runtime errors. This blog entry has not been intended in any way to be a definitive work on IoC, but rather as something to spur thought about why we program to interfaces and some ways to reach the intended value of the practice instead of having it just complicate your code.  I hope that it helps somebody begin or continue a journey away from being a "Cargo Cult Programmer".

    Read the article

  • Creating a Document Library with Content Type in code

    - by David Jacobus
    Originally posted on: http://geekswithblogs.net/djacobus/archive/2013/10/15/154360.aspxIn the past, I have shown how to create a list content type and add the content type to a list in code.  As a Developer, many of the artifacts which we create are widgets which have a List or Document Library as the back end.   We need to be able to create our applications (Web Part, etc.) without having the user involved except to enter the list item data.  Today, I will show you how to do the same with a document library.    A summary of what we will do is as follows:   1.   Create an Empty SharePoint Project in Visual Studio 2.   Add a Code Folder in the solution and Drag and Drop Utilities and Extensions Libraries to the solution 3.   Create a new Feature and add and event receiver  all the code will be in the event receiver 4.   Add the fields which will extend the built-in Document content type 5.   If the Content Type does not exist, Create it 6.   If the Document Library does not exist, Create it with the new Content Type inherited from the Document Content Type 7.   Delete the Document Content Type from the Library (as we have a new one which inherited from it) 8.   Add the fields which we want to be visible from the fields added to the new Content Type   Here we go:   Create an Empty SharePoint Project in Visual Studio      Add a Code Folder in the solution and Drag and Drop Utilities and Extensions Libraries to the solution       The Utilities and Extensions Library will be part of this project which I will provide a download link at the end of this post.  Drag and drop them into your project.  If Dragged and Dropped from windows explorer you will need to show all files and then include them in your project.  Change the Namespace to agree with your project.   Create a new Feature and add and event receiver  all the code will be in the event receiver.  Here We added a new Feature called “CreateDocLib”  and then right click to add an Event Receiver All of our code will be in this Event Receiver.  For this Demo I will only be using the Feature Activated Event.      From this point on we will be looking at code!    We are adding two constants for use columGroup (How we want SharePoint to Group them, usually Company Name) and ctName(ContentType Name)  using System; using System.Runtime.InteropServices; using System.Security.Permissions; using Microsoft.SharePoint; namespace CreateDocLib.Features.CreateDocLib { /// <summary> /// This class handles events raised during feature activation, deactivation, installation, uninstallation, and upgrade. /// </summary> /// <remarks> /// The GUID attached to this class may be used during packaging and should not be modified. /// </remarks> [Guid("56e6897c-97c4-41ac-bc5b-5cd2c04f2dd1")] public class CreateDocLibEventReceiver : SPFeatureReceiver { const string columnGroup = "DJ"; const string ctName = "DJDocLib"; } }     Here we are creating the Feature Activated event.   Adding the new fields (Site Columns) ,  Testing if the Content Type Exists, if not adding it.  Testing if the document Library exists, if not adding it.   #region DocLib public override void FeatureActivated(SPFeatureReceiverProperties properties) { using (SPWeb spWeb = properties.GetWeb() as SPWeb) { //add the fields addFields(spWeb); //add content type SPContentType testCT = spWeb.ContentTypes[ctName]; // we will not create the content type if it exists if (testCT == null) { //the content type does not exist add it addContentType(spWeb, ctName); } if ((spWeb.Lists.TryGetList("MyDocuments") == null)) { //create the list if it dosen't to exist CreateDocLib(spWeb); } } } #endregion The addFields method uses the utilities library to add site columns to the site. We can add as many fields within this method as we like. Here we are adding one for demonstration purposes. Icon as a Url type.  public void addFields(SPWeb spWeb) { Utilities.addField(spWeb, "Icon", SPFieldType.URL, false, columnGroup); }The addContentType method add the new Content Type to the site Content Types. We have already checked to see that it does not exist. In addition, here is where we add the linkages from our site columns previously created to our new Content Type   private static void addContentType(SPWeb spWeb, string name) { SPContentType myContentType = new SPContentType(spWeb.ContentTypes["Document"], spWeb.ContentTypes, name) { Group = columnGroup }; spWeb.ContentTypes.Add(myContentType); addContentTypeLinkages(spWeb, myContentType); myContentType.Update(); } Here we are adding just one linkage as we only have one additional field in our Content Type public static void addContentTypeLinkages(SPWeb spWeb, SPContentType ct) { Utilities.addContentTypeLink(spWeb, "Icon", ct); } Next we add the logic to create our new Document Library, which we have already checked to see if it exists.  We create the document library and turn on content types.  Add the new content type and then delete the old “Document” content types.   private void CreateDocLib(SPWeb web) { using (var site = new SPSite(web.Url)) { var web1 = site.RootWeb; var listId = web1.Lists.Add("MyDocuments", string.Empty, SPListTemplateType.DocumentLibrary); var lib = web1.Lists[listId] as SPDocumentLibrary; lib.ContentTypesEnabled = true; var docType = web.ContentTypes[ctName]; lib.ContentTypes.Add(docType); lib.ContentTypes.Delete(lib.ContentTypes["Document"].Id); lib.Update(); AddLibrarySettings(web1, lib); } }  Finally, we set some document library settings on our new document library with the AddLibrarySettings method. We then ensure that the new site column is visible when viewed in the browser.  private void AddLibrarySettings(SPWeb web, SPDocumentLibrary lib) { lib.OnQuickLaunch = true; lib.ForceCheckout = true; lib.EnableVersioning = true; lib.MajorVersionLimit = 5; lib.EnableMinorVersions = true; lib.MajorWithMinorVersionsLimit = 5; lib.Update(); var view = lib.DefaultView; view.ViewFields.Add("Icon"); view.Update(); } Okay, what's cool here: In a few lines of code, we have created site columns, A content Type, a document library. As a developer, I use this functionality all the time. For instance, I could now just add a web part to this same solutionwhich uses this document Library. I love SharePoint! Here is the complete solution: Create Document Library Code

    Read the article

  • Visual Studio 2010 Productivity Tips and Tricks&ndash;Part 1: Extensions

    - by ToStringTheory
    I don’t know about you, but when it comes to development, I prefer my environment to be as free of clutter as possible.  It may surprise you to know that I have tried ReSharper, and did not like it, for the reason that I stated above.  In my opinion, it had too much clutter.  Don’t get me wrong, there were a couple of features that I did like about it (inversion of if blocks, code feedback), but for the most part, I actually felt that it was slowing me down. Introduction Another large factor besides intrusiveness/speed in my choice to dislike ReSharper would probably be that I have become comfortable with my current setup and extensions.  I believe I have a good collection, and am quite happy with what I can accomplish in a short amount of time.  I figured that I would share some of my tips/findings regarding Visual Studio productivity here, and see what you had to say. The first section of things that I would like to cover, are Visual Studio Extensions.  In case you have been living under a rock for the past several years, Extensions are available under the Tools menu in Visual Studio: The extension manager enables integrated access to the Microsoft Visual Studio Gallery online with access to a few thousand different extensions.  I have tried many extensions, but for reasons of lack reliability, usability, or features, have uninstalled almost all of them.  However, I have come across several that I find I can not do without anymore: NuGet Package Manager (Microsoft) Perspectives (Adam Driscoll) Productivity Power Tools (Microsoft) Web Essentials (Mads Kristensen) Extensions NuGet Package Manager To be honest, I debated on whether or not to put this in here.  Most people seem to have it, however, there was a time when I didn’t, and was always confused when blogs/posts would say to right click and “Add Package Reference…” which with one of the latest updates is now “Manage NuGet Packages”.  So, if you haven’t downloaded the NuGet Package Manager yet, or don’t know what it is, I would highly suggest downloading it now! Features Simply put, the NuGet Package Manager gives you a GUI and command line to access different libraries that have been uploaded to NuGet. Some of its features include: Ability to search NuGet for packages via the GUI, with information in the detail bar on the right. Quick access to see what packages are in a solution, and what packages have updates available, with easy 1-click updating. If you download a package that requires references to work on other NuGet packages, they will be downloaded and referenced automatically. Productivity Tip If you use any type of source control in Visual Studio as well as using NuGet packages, be sure to right-click on the solution and click "Enable NuGet Package Restore". What this does is add a NuGet package to the solution so that it will be checked in along side your solution, as well as automatically grab packages from NuGet on build if needed. This is an extremely simple system to use to manage your package references, instead of having to manually go into TFS and add the Packages folder. Perspectives I can't stand developing with just one monitor. Especially if it comes to debugging. The great thing about Visual Studio 2010, is that all of the panels and windows are floatable, and can dock to other screens. The only bad thing is, I don't use the same toolset with everything that I am doing. By this, I mean that I don't use all of the same windows for debugging a web application, as I do for coding a WPF application. Only thing is, Visual Studio doesn't save the screen positions for all of the undocked windows. So, I got curious one day and decided to check and see if there was an extension to help out. This is where I found Perspectives. Features Perspectives gives you the ability to configure window positions across any or your monitors, and then to save the positions in a profile. Perspectives offers a Panel to manage different presets/favorites, and a toolbar to add to the toolbars at the top of Visual Studio. Ability to 'Favorite' a profile to add it to the perspectives toolbar. Productivity Tip Take the time to setup profiles for each of your scenarios - debugging web/winforms/xaml, coding, maintenance, etc. Try to remember to use the profiles for a few days, and at the end of a week, you may find that your productivity was never better. Productivity Power Tools Ah, the Productivity Power Tools... Quite possibly one of my most used extensions, if not my most used. The tool pack gives you a variety of enhancements ranging from key shortcuts, interface tweaks, and completely new features to Visual Studio 2010. Features I don't want to bore you with all of the features here, so here are my favorite: Quick Find - Unobtrusive search box in upper-right corner of the code window. Great for searching in general, especially in a file. Solution Navigator - The 'Solution Explorer' on steroids. Easy to search for files, see defined members/properties/methods in files, and my favorite feature is the 'set as root' option. Updated 'Add Reference...' Dialog - This is probably my favorite enhancement period... The 'Add Reference...' dialog redone in a manner that resembles the Extension/Package managers. I especially love the ability to search through all of the references. "Ctrl - Click" for Definition - I am still getting used to this as I usually try to use my keyboard for everything, but I love the ability to hold Ctrl and turn property/methods/variables into hyperlinks, that you click on to see their definitions. Great for travelling down a rabbit hole in an application to research problems. While there are other commands/utilities, I find these to be the ones that I lean on the most for the usefulness. Web Essentials If you have do any type of web development in ASP .Net, ASP .Net MVC, even HTML, I highly suggest grabbing the Web Essentials right NOW! This extension alone is great for productivity in web development, and greatly decreases my development time on new features. Features Some of its best features include: CSS Previews - I say 'previews' because of the multiple kinds of previews in CSS that you get font-family, color, background/background-image previews. This is great for just tweaking UI slightly in different ways and seeing how they look in the CSS window at a glance. Live Preview - One word - awesome! This goes well with my multi-monitor setup. I put the site on one monitor in a Live Preview panel, and then as I make changes to CSS/cshtml/aspx/html, the preview window will update with each save/build automatically. For CSS, you can even turn on live-update, so as you are tweaking CSS, the style changes in real time. Great for tweaking colors or font-sizes. Outlining - Small, but I like to be able to collapse regions/declarations that are in the way of new work, or are just distracting. Commenting Shortcuts - I don't know why it wasn't included by default, but it is nice to have the key shortcuts for commenting working in the CSS editor as well. Productivity Tip When working on a site, hit CTRL-ALT-ENTER to launch the Live Preview window. Dock it to another monitor. When you make changes to the document/css, just save and glance at the other monitor. No need to alt tab, then alt tab before continuing editing. Conclusion These extensions are only the most useful and least intrusive - ones that I use every day. The great thing about Visual Studio 2010 is the extensibility options that it gives developers to utilize. Have an extension that you use that isn't intrusive, but isn't listed here? Please, feel free to comment. I love trying new things, and am always looking for new additions to my toolset of the most useful. Finally, please keep an eye out for Part 2 on key shortcuts in Visual Studio. Also, if you are visiting my site (http://tostringtheory.com || http://geekswithblogs.net/tostringtheory) from an actual browser and not a feed, please let me know what you think of the new styling!

    Read the article

  • Applications: The Mathematics of Movement, Part 2

    - by TechTwaddle
    In part 1 of this series we saw how we can make the marble move towards the click point, with a fixed speed. In this post we’ll see, first, how to get rid of Atan2(), sine() and cosine() in our calculations, and, second, reducing the speed of the marble as it approaches the destination, so it looks like the marble is easing into it’s final position. As I mentioned in one of the previous posts, this is achieved by making the speed of the marble a function of the distance between the marble and the destination point. Getting rid of Atan2(), sine() and cosine() Ok, to be fair we are not exactly getting rid of these trigonometric functions, rather, replacing one form with another. So instead of writing sin(?), we write y/length. You see the point. So instead of using the trig functions as below, double x = destX - marble1.x; double y = destY - marble1.y; //distance between destination and current position, before updating marble position distanceSqrd = x * x + y * y; double angle = Math.Atan2(y, x); //Cos and Sin give us the unit vector, 6 is the value we use to magnify the unit vector along the same direction incrX = speed * Math.Cos(angle); incrY = speed * Math.Sin(angle); marble1.x += incrX; marble1.y += incrY; we use the following, double x = destX - marble1.x; double y = destY - marble1.y; //distance between destination and marble (before updating marble position) lengthSqrd = x * x + y * y; length = Math.Sqrt(lengthSqrd); //unit vector along the same direction as vector(x, y) unitX = x / length; unitY = y / length; //update marble position incrX = speed * unitX; incrY = speed * unitY; marble1.x += incrX; marble1.y += incrY; so we replaced cos(?) with x/length and sin(?) with y/length. The result is the same.   Adding oomph to the way it moves In the last post we had the speed of the marble fixed at 6, double speed = 6; to make the marble decelerate as it moves, we have to keep updating the speed of the marble in every frame such that the speed is calculated as a function of the length. So we may have, speed = length/12; ‘length’ keeps decreasing as the marble moves and so does speed. The Form1_MouseUp() function remains the same as before, here is the UpdatePosition() method, private void UpdatePosition() {     double incrX = 0, incrY = 0;     double lengthSqrd = 0, length = 0, lengthSqrdNew = 0;     double unitX = 0, unitY = 0;     double speed = 0;     double x = destX - marble1.x;     double y = destY - marble1.y;     //distance between destination and marble (before updating marble position)     lengthSqrd = x * x + y * y;     length = Math.Sqrt(lengthSqrd);     //unit vector along the same direction as vector(x, y)     unitX = x / length;     unitY = y / length;     //speed as a function of length     speed = length / 12;     //update marble position     incrX = speed * unitX;     incrY = speed * unitY;     marble1.x += incrX;     marble1.y += incrY;     //check for bounds     if ((int)marble1.x < MinX + marbleWidth / 2)     {         marble1.x = MinX + marbleWidth / 2;     }     else if ((int)marble1.x > (MaxX - marbleWidth / 2))     {         marble1.x = MaxX - marbleWidth / 2;     }     if ((int)marble1.y < MinY + marbleHeight / 2)     {         marble1.y = MinY + marbleHeight / 2;     }     else if ((int)marble1.y > (MaxY - marbleHeight / 2))     {         marble1.y = MaxY - marbleHeight / 2;     }     //distance between destination and marble (after updating marble position)     x = destX - (marble1.x);     y = destY - (marble1.y);     lengthSqrdNew = x * x + y * y;     /*      * End Condition:      * 1. If there is not much difference between lengthSqrd and lengthSqrdNew      * 2. If the marble has moved more than or equal to a distance of totLenToTravel (see Form1_MouseUp)      */     x = startPosX - marble1.x;     y = startPosY - marble1.y;     double totLenTraveledSqrd = x * x + y * y;     if ((int)totLenTraveledSqrd >= (int)totLenToTravelSqrd)     {         System.Console.WriteLine("Stopping because Total Len has been traveled");         timer1.Enabled = false;     }     else if (Math.Abs((int)lengthSqrd - (int)lengthSqrdNew) < 4)     {         System.Console.WriteLine("Stopping because no change in Old and New");         timer1.Enabled = false;     } } A point to note here is that, in this implementation, the marble never stops because it travelled a distance of totLenToTravelSqrd (first if condition). This happens because speed is a function of the length. During the final few frames length becomes very small and so does speed; and so the amount by which the marble shifts is quite small, and the second if condition always hits true first. I’ll end this series with a third post. In part 3 we will cover two things, one, when the user clicks, the marble keeps moving in that direction, rebounding off the screen edges and keeps moving forever. Two, when the user clicks on the screen, the marble moves towards it, with it’s speed reducing by every frame. It doesn’t come to a halt when the destination point is reached, instead, it continues to move, rebounds off the screen edges and slowly comes to halt. The amount of time that the marble keeps moving depends on how far the user clicks from the marble. I had mentioned this second situation here. Finally, here’s a video of this program running,

    Read the article

  • We've completed the first iteration

    - by CliveT
    There are a lot of features in C# that are implemented by the compiler and not by the underlying platform. One such feature is a lambda expression. Since local variables cannot be accessed once the current method activation finishes, the compiler has to go out of its way to generate a new class which acts as a home for any variable whose lifetime needs to be extended past the activation of the procedure. Take the following example:     Random generator = new Random();     Func func = () = generator.Next(10); In this case, the compiler generates a new class called c_DisplayClass1 which is marked with the CompilerGenerated attribute. [CompilerGenerated] private sealed class c__DisplayClass1 {     // Fields     public Random generator;     // Methods     public int b__0()     {         return this.generator.Next(10);     } } Two quick comments on this: (i)    A display was the means that compilers for languages like Algol recorded the various lexical contours of the nested procedure activations on the stack. I imagine that this is what has led to the name. (ii)    It is a shame that the same attribute is used to mark all compiler generated classes as it makes it hard to figure out what they are being used for. Indeed, you could imagine optimisations that the runtime could perform if it knew that classes corresponded to certain high level concepts. We can see that the local variable generator has been turned into a field in the class, and the body of the lambda expression has been turned into a method of the new class. The code that builds the Func object simply constructs an instance of this class and initialises the fields to their initial values.     c__DisplayClass1 class2 = new c__DisplayClass1();     class2.generator = new Random();     Func func = new Func(class2.b__0); Reflector already contains code to spot this pattern of code and reproduce the form containing the lambda expression, so this is example is correctly decompiled. The use of compiler generated code is even more spectacular in the case of iterators. C# introduced the idea of a method that could automatically store its state between calls, so that it can pick up where it left off. The code can express the logical flow with yield return and yield break denoting places where the method should return a particular value and be prepared to resume.         {             yield return 1;             yield return 2;             yield return 3;         } Of course, there was already a .NET pattern for expressing the idea of returning a sequence of values with the computation proceeding lazily (in the sense that the work for the next value is executed on demand). This is expressed by the IEnumerable interface with its Current property for fetching the current value and the MoveNext method for forcing the computation of the next value. The sequence is terminated when this method returns false. The C# compiler links these two ideas together so that an IEnumerator returning method using the yield keyword causes the compiler to produce the implementation of an Iterator. Take the following piece of code.         IEnumerable GetItems()         {             yield return 1;             yield return 2;             yield return 3;         } The compiler implements this by defining a new class that implements a state machine. This has an integer state that records which yield point we should go to if we are resumed. It also has a field that records the Current value of the enumerator and a field for recording the thread. This latter value is used for optimising the creation of iterator instances. [CompilerGenerated] private sealed class d__0 : IEnumerable, IEnumerable, IEnumerator, IEnumerator, IDisposable {     // Fields     private int 1__state;     private int 2__current;     public Program 4__this;     private int l__initialThreadId; The body gets converted into the code to construct and initialize this new class. private IEnumerable GetItems() {     d__0 d__ = new d__0(-2);     d__.4__this = this;     return d__; } When the class is constructed we set the state, which was passed through as -2 and the current thread. public d__0(int 1__state) {     this.1__state = 1__state;     this.l__initialThreadId = Thread.CurrentThread.ManagedThreadId; } The state needs to be set to 0 to represent a valid enumerator and this is done in the GetEnumerator method which optimises for the usual case where the returned enumerator is only used once. IEnumerator IEnumerable.GetEnumerator() {     if ((Thread.CurrentThread.ManagedThreadId == this.l__initialThreadId)               && (this.1__state == -2))     {         this.1__state = 0;         return this;     } The state machine itself is implemented inside the MoveNext method. private bool MoveNext() {     switch (this.1__state)     {         case 0:             this.1__state = -1;             this.2__current = 1;             this.1__state = 1;             return true;         case 1:             this.1__state = -1;             this.2__current = 2;             this.1__state = 2;             return true;         case 2:             this.1__state = -1;             this.2__current = 3;             this.1__state = 3;             return true;         case 3:             this.1__state = -1;             break;     }     return false; } At each stage, the current value of the state is used to determine how far we got, and then we generate the next value which we return after recording the next state. Finally we return false from the MoveNext to signify the end of the sequence. Of course, that example was really simple. The original method body didn't have any local variables. Any local variables need to live between the calls to MoveNext and so they need to be transformed into fields in much the same way that we did in the case of the lambda expression. More complicated MoveNext methods are required to deal with resources that need to be disposed when the iterator finishes, and sometimes the compiler uses a temporary variable to hold the return value. Why all of this explanation? We've implemented the de-compilation of iterators in the current EAP version of Reflector (7). This contrasts with previous version where all you could do was look at the MoveNext method and try to figure out the control flow. There's a fair amount of things we have to do. We have to spot the use of a CompilerGenerated class which implements the Enumerator pattern. We need to go to the class and figure out the fields corresponding to the local variables. We then need to go to the MoveNext method and try to break it into the various possible states and spot the state transitions. We can then take these pieces and put them back together into an object model that uses yield return to show the transition points. After that Reflector can carry on optimising using its usual optimisations. The pattern matching is currently a little too sensitive to changes in the code generation, and we only do a limited analysis of the MoveNext method to determine use of the compiler generated fields. In some ways, it is a pity that iterators are compiled away and there is no metadata that reflects the original intent. Without it, we are always going to dependent on our knowledge of the compiler's implementation. For example, we have noticed that the Async CTP changes the way that iterators are code generated, so we'll have to do some more work to support that. However, with that warning in place, we seem to do a reasonable job of decompiling the iterators that are built into the framework. Hopefully, the EAP will give us a chance to find examples where we don't spot the pattern correctly or regenerate the wrong code, and we can improve things. Please give it a go, and report any problems.

    Read the article

  • Per-pixel displacement mapping GLSL

    - by Chris
    Im trying to implement a per-pixel displacement shader in GLSL. I read through several papers and "tutorials" I found and ended up with trying to implement the approach NVIDIA used in their Cascade Demo (http://www.slideshare.net/icastano/cascades-demo-secrets) starting at Slide 82. At the moment I am completly stuck with following problem: When I am far away the displacement seems to work. But as more I move closer to my surface, the texture gets bent in x-axis and somehow it looks like there is a little bent in general in one direction. EDIT: I added a video: click I added some screen to illustrate the problem: Well I tried lots of things already and I am starting to get a bit frustrated as my ideas run out. I added my full VS and FS code: VS: #version 400 layout(location = 0) in vec3 IN_VS_Position; layout(location = 1) in vec3 IN_VS_Normal; layout(location = 2) in vec2 IN_VS_Texcoord; layout(location = 3) in vec3 IN_VS_Tangent; layout(location = 4) in vec3 IN_VS_BiTangent; uniform vec3 uLightPos; uniform vec3 uCameraDirection; uniform mat4 uViewProjection; uniform mat4 uModel; uniform mat4 uView; uniform mat3 uNormalMatrix; out vec2 IN_FS_Texcoord; out vec3 IN_FS_CameraDir_Tangent; out vec3 IN_FS_LightDir_Tangent; void main( void ) { IN_FS_Texcoord = IN_VS_Texcoord; vec4 posObject = uModel * vec4(IN_VS_Position, 1.0); vec3 normalObject = (uModel * vec4(IN_VS_Normal, 0.0)).xyz; vec3 tangentObject = (uModel * vec4(IN_VS_Tangent, 0.0)).xyz; //vec3 binormalObject = (uModel * vec4(IN_VS_BiTangent, 0.0)).xyz; vec3 binormalObject = normalize(cross(tangentObject, normalObject)); // uCameraDirection is the camera position, just bad named vec3 fvViewDirection = normalize( uCameraDirection - posObject.xyz); vec3 fvLightDirection = normalize( uLightPos.xyz - posObject.xyz ); IN_FS_CameraDir_Tangent.x = dot( tangentObject, fvViewDirection ); IN_FS_CameraDir_Tangent.y = dot( binormalObject, fvViewDirection ); IN_FS_CameraDir_Tangent.z = dot( normalObject, fvViewDirection ); IN_FS_LightDir_Tangent.x = dot( tangentObject, fvLightDirection ); IN_FS_LightDir_Tangent.y = dot( binormalObject, fvLightDirection ); IN_FS_LightDir_Tangent.z = dot( normalObject, fvLightDirection ); gl_Position = (uViewProjection*uModel) * vec4(IN_VS_Position, 1.0); } The VS just builds the TBN matrix, from incoming normal, tangent and binormal in world space. Calculates the light and eye direction in worldspace. And finally transforms the light and eye direction into tangent space. FS: #version 400 // uniforms uniform Light { vec4 fvDiffuse; vec4 fvAmbient; vec4 fvSpecular; }; uniform Material { vec4 diffuse; vec4 ambient; vec4 specular; vec4 emissive; float fSpecularPower; float shininessStrength; }; uniform sampler2D colorSampler; uniform sampler2D normalMapSampler; uniform sampler2D heightMapSampler; in vec2 IN_FS_Texcoord; in vec3 IN_FS_CameraDir_Tangent; in vec3 IN_FS_LightDir_Tangent; out vec4 color; vec2 TraceRay(in float height, in vec2 coords, in vec3 dir, in float mipmap){ vec2 NewCoords = coords; vec2 dUV = - dir.xy * height * 0.08; float SearchHeight = 1.0; float prev_hits = 0.0; float hit_h = 0.0; for(int i=0;i<10;i++){ SearchHeight -= 0.1; NewCoords += dUV; float CurrentHeight = textureLod(heightMapSampler,NewCoords.xy, mipmap).r; float first_hit = clamp((CurrentHeight - SearchHeight - prev_hits) * 499999.0,0.0,1.0); hit_h += first_hit * SearchHeight; prev_hits += first_hit; } NewCoords = coords + dUV * (1.0-hit_h) * 10.0f - dUV; vec2 Temp = NewCoords; SearchHeight = hit_h+0.1; float Start = SearchHeight; dUV *= 0.2; prev_hits = 0.0; hit_h = 0.0; for(int i=0;i<5;i++){ SearchHeight -= 0.02; NewCoords += dUV; float CurrentHeight = textureLod(heightMapSampler,NewCoords.xy, mipmap).r; float first_hit = clamp((CurrentHeight - SearchHeight - prev_hits) * 499999.0,0.0,1.0); hit_h += first_hit * SearchHeight; prev_hits += first_hit; } NewCoords = Temp + dUV * (Start - hit_h) * 50.0f; return NewCoords; } void main( void ) { vec3 fvLightDirection = normalize( IN_FS_LightDir_Tangent ); vec3 fvViewDirection = normalize( IN_FS_CameraDir_Tangent ); float mipmap = 0; vec2 NewCoord = TraceRay(0.1,IN_FS_Texcoord,fvViewDirection,mipmap); //vec2 ddx = dFdx(NewCoord); //vec2 ddy = dFdy(NewCoord); vec3 BumpMapNormal = textureLod(normalMapSampler, NewCoord.xy, mipmap).xyz; BumpMapNormal = normalize(2.0 * BumpMapNormal - vec3(1.0, 1.0, 1.0)); vec3 fvNormal = BumpMapNormal; float fNDotL = dot( fvNormal, fvLightDirection ); vec3 fvReflection = normalize( ( ( 2.0 * fvNormal ) * fNDotL ) - fvLightDirection ); float fRDotV = max( 0.0, dot( fvReflection, fvViewDirection ) ); vec4 fvBaseColor = textureLod( colorSampler, NewCoord.xy,mipmap); vec4 fvTotalAmbient = fvAmbient * fvBaseColor; vec4 fvTotalDiffuse = fvDiffuse * fNDotL * fvBaseColor; vec4 fvTotalSpecular = fvSpecular * ( pow( fRDotV, fSpecularPower ) ); color = ( fvTotalAmbient + (fvTotalDiffuse + fvTotalSpecular) ); } The FS implements the displacement technique in TraceRay method, while always using mipmap level 0. Most of the code is from NVIDIA sample and another paper I found on the web, so I guess there cannot be much wrong in here. At the end it uses the modified UV coords for getting the displaced normal from the normal map and the color from the color map. I looking forward for some ideas. Thanks in advance! Edit: Here is the code loading the heightmap: glTexImage2D(GL_TEXTURE_2D, 0, GL_RGBA, mWidth, mHeight, 0, GL_RGBA, GL_UNSIGNED_BYTE, mImageData); glGenerateMipmap(GL_TEXTURE_2D); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MAG_FILTER, GL_LINEAR_MIPMAP_LINEAR); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MIN_FILTER, GL_LINEAR_MIPMAP_LINEAR); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_S, GL_REPEAT); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_T, GL_REPEAT); Maybe something wrong in here?

    Read the article

  • Google Analytics on Android

    - by pjv
    There is a specific and official analytics SDK for native Android apps (note that I'm not talking about webpages in apps on a phone). This library basically sends pages and events to Google Analytics and you can view your analytics in exactly the same dashboard as for websites. Since my background is apps rather than websites, and since a lot of the Google Analytics terminology seems particularly inapplicable to a native app, I need some pointers. Please discuss my remarks, provide some clarification where you think I'm off-track, and above all share good experiences! 1. Page Views Pages mostly can match different Activities (and Dialogs) being displayed. Activities can be visible behind non-full-screen Activities however, though only the top-level Activity can be interacted. This sort-off clashes with a "(page) view". You'd also want at least one page view for each visit and therefore put one page view tracker in the Application class. However this does not constitute a window or sorts. Usually an Activity will open at the same time, so the time spent on that page will have been 0. This will influence your "time spent" statistics. How are these counted anyway? Moreover, there is a loose coupling between the Activities, by means of Intents. A user can, much like on any website, step in at any Activity, although usually this then concerns resuming the application where he left off. This makes that the hierarchy of Activities usually is very flat. And since there are no url's involved. What meaning would using slashes in page titles have, such as "/Home"? All pages would appear on an equal level in the reports, so no content drilldown. Non-unique page views seem to be counted as some kind of indicator of successfulness: how often does the visitor revisit the page. When the user rotates the screen however usually an Activity resumes again, thus making it a new page view. This happens a lot. Maybe a well-thought-through placement of the call might solve this, or placing several, I'm not sure. How to deal with Page Views? 2. Events I'd say there are two sorts: A user event Something that happened, usually as an indirect consequence of the above. The latter particularly is giving me headaches. First of all, many events aren't written in code any more, but pieced logically together by means of Intents. This means that there is no place to put the analytics call. You'd either have to give up this advantage and start doing it the old-fashioned way in favor of good analytics, or, just be missing some events. Secondly, as a developer you're not so much interested in when a user clicks a button, but if the action that should have been performed really was performed and what the result was. There seems to be no clear way to get resulting data into Google Analytics (what's up with the integers? I want to put in Strings!). The same that applies to the flat pages hierarchy, also goes for the event categories. You could do "vertical" categories (topically, that is), but some code is shared "horizontally" and the tracking will be equally shared. Just as with the Intents mechanism, inheritance makes it hard for you to put the tracking in the right places at all times. And I can't really imagine "horizontal" categories. Unless you start making really small categories, such as all the items form the same menu in one category, I have a hard time grasping the concept. Finally, how do you deal with cancelling? Usually you both have an explicit cancel mechanism by ways of a button, as well as the implicit cancel when the "back"-button is pressed to leave the activity and there were no changes. The latter also applies to "saves", when the back button is pressed and there ARE changes. How are you consequently going to catch all these if not by doing all the "back"-button work yourself? How to deal with events? 3. Goals For goal types I have choice of: URL Destination, Time on Site, and Pages/Visit. Most apps don't have a funnel that leads the user to some "registration done" or "order placed" page. Apps have either already been bought (in which case you want to stimulate the user to love your app, so that he might bring on new buyers) or are paid for by in-app ads. So URL Destination is not a very important goal. Time on Site also seems troublesome. First, I have some doubt on how this would be measured. Second, I don't necessarily want my user to spend a lot of time in my already paid app, just be active and content. Equivalently, why not mention how frequent a user uses your app? Regarding Pages/Visit I already mentioned how screen orientation changes blow up the page view numbers. In an app I'd be most interested in events/visit to measure the user's involvement/activity. If he's intensively using the app then he must be loving it right? Furthermore, I also have some small funnels (that do not lead to conversion though) that I want to see streamlined. In my mind those funnels would end in events rather than page views but that seems not to be possible. I could also measure clickthroughs on in-app ads, but then I'd need to track those as Page Views rather than Events, in view of "URL Destination". What are smart goals for apps and how can you fit them on top of Analytics? 4. Optimisation Is there a smart way to manually do what "Website Optimiser" does for websites? Most importantly, how would I track different landing page designs? 5. Traffic Sources Referrals deal with installation time referrals, if you're smart enough to get them included. But perhaps I'd also want to get some data which third-party app sends users to my app to perform some actions (this app interoperability is possible via Intents). Many of the terminologies related to "Traffic Sources" seem totally meaningless and there is no possibility of connecting in AdSense. What are smart uses of this data? 6. Visitors Of the "Browser capabilities", "Network Properties" and "Mobile" tabs, many things are pointless as they have no influence on / relation with my mostly offline app that won't use flash anyway. Only if you drill down far enough, can you get to OS versions, which do matter a lot. I even forgot where you could check what exact Android devices visited. What are smart uses of this data? How can you make the relevant info more prominent? 7. Other No in-page analytics. I have to register my app as a web-url (What!?)?

    Read the article

  • Who could ask for more with LESS CSS? (Part 3 of 3&ndash;Clrizr)

    - by ToString(theory);
    Welcome back!  In the first two posts in this series, I covered some of the awesome features in CSS precompilers such as SASS and LESS, as well as how to get an initial project setup up and running in ASP.Net MVC 4. In this post, I will cover an actual advanced example of using LESS in a project, and show some of the great productivity features we gain from its usage. Introduction In the first post, I mentioned two subjects that I will be using in this example – constants, and color functions.  I’ve always enjoyed using online color scheme utilities such as Adobe Kuler or Color Scheme Designer to come up with a scheme based off of one primary color.  Using these tools, and requesting a complementary scheme you can get a couple of shades of your primary color, and a couple of shades of a complementary/accent color to display. Because there is no way in regular css to do color operations or store variables, there was no way to accomplish something like defining a primary color, and have a site theme cascade off of that.  However with tools such as LESS, that impossibility becomes a reality!  So, if you haven’t guessed it by now, this post is on the creation of a plugin/module/less file to drop into your project, plugin one color, and have your primary theme cascade from it.  I only went through the trouble of creating a module for getting Complementary colors.  However, it wouldn’t be too much trouble to go through other options such as Triad or Monochromatic to get a module that you could use off of that. Step 1 – Analysis I decided to mimic Adobe Kuler’s Complementary theme algorithm as I liked its simplicity and aesthetics.  Color Scheme Designer is great, but I do believe it can give you too many color options, which can lead to chaos and overload.  The first thing I had to check was if the complementary values for the color schemes were actually hues rotated by 180 degrees at all times – they aren’t.  Apparently Adobe applies some variance to the complementary colors to get colors that are actually more aesthetically appealing to users.  So, I opened up Excel and began to plot complementary hues based on rotation in increments of 10: Long story short, I completed the same calculations for Hue, Saturation, and Lightness.  For Hue, I only had to record the Complementary hue values, however for saturation and lightness, I had to record the values for ALL of the shades.  Since the functions were too complicated to put into LESS since they aren’t constant/linear, but rather interval functions, I instead opted to extrapolate the HSL values using the trendline function for each major interval, onto intervals of spacing 1. For example, using the hue extraction, I got the following values: Interval Function 0-60 60-140 140-270 270-360 Saturation and Lightness were much worse, but in the end, I finally had functions for all of the intervals, and then went the route of just grabbing each shades value in intervals of 1.  Step 2 – Mapping I declared variable names for each of these sections as something that shouldn’t ever conflict with a variable someone would define in their own file.  After I had each of the values, I extracted the values and put them into files of their own for hue variables, saturation variables, and lightness variables…  Example: /*HUE CONVERSIONS*/@clrizr-hue-source-0deg: 133.43;@clrizr-hue-source-1deg: 135.601;@clrizr-hue-source-2deg: 137.772;@clrizr-hue-source-3deg: 139.943;@clrizr-hue-source-4deg: 142.114;.../*SATURATION CONVERSIONS*/@clrizr-saturation-s2SV0px: 0;@clrizr-saturation-s2SV1px: 0;@clrizr-saturation-s2SV2px: 0;@clrizr-saturation-s2SV3px: 0;@clrizr-saturation-s2SV4px: 0;.../*LIGHTNESS CONVERSIONS*/@clrizr-lightness-s2LV0px: 30;@clrizr-lightness-s2LV1px: 31;@clrizr-lightness-s2LV2px: 32;@clrizr-lightness-s2LV3px: 33;@clrizr-lightness-s2LV4px: 34;...   In the end, I have 973 lines of mapping/conversion from source HSL to shade HSL for two extra primary shades, and two complementary shades. The last bit of the work was the file to compose each of the shades from these mappings. Step 3 – Clrizr Mapper The final step was the hardest to overcome as I was still trying to understand LESS to its fullest extent.  Imports As mentioned previously, I had separated the HSL mappings into different files, so the first necessary step is to import those for use into the Clrizr plugin: @import url("hue.less");@import url("saturation.less");@import url("lightness.less"); Extract Component Values For Each Shade Next, I extracted the necessary information for each shade HSL before shade composition: @clrizr-input-saturation: 1px+floor(saturation(@clrizr-input))-1;@clrizr-input-lightness: 1px+floor(lightness(@clrizr-input))-1; @clrizr-complementary-hue: formatstring("clrizr-hue-source-{0}", ceil(hue(@clrizr-input))); @clrizr-primary-2-saturation: formatstring("clrizr-saturation-s2SV{0}",@clrizr-input-saturation);@clrizr-primary-1-saturation: formatstring("clrizr-saturation-s1SV{0}",@clrizr-input-saturation);@clrizr-complementary-1-saturation: formatstring("clrizr-saturation-c1SV{0}",@clrizr-input-saturation); @clrizr-primary-2-lightness: formatstring("clrizr-lightness-s2LV{0}",@clrizr-input-lightness);@clrizr-primary-1-lightness: formatstring("clrizr-lightness-s1LV{0}",@clrizr-input-lightness);@clrizr-complementary-1-lightness: formatstring("clrizr-lightness-c1LV{0}",@clrizr-input-lightness); Here, you can see a couple of odd things…  On the first line, I am using operations to add units to the saturation and lightness.  This is due to some limitations in the operations that would give me saturation or lightness in %, which can’t be in a variable name.  So, I use first add 1px to it, which casts the result of the following functions as px instead of %, and then at the end, I remove that pixel.  You can also see here the formatstring method which is exactly what it sounds like – something like String.Format(string str, params object[] obj). Get Primary & Complementary Shades Now that I have components for each of the different shades, I can now compose them into each of their pieces.  For this, I use the @@ operator which will look for a variable with the name specified in a string, and then call that variable: @clrizr-primary-2: hsl(hue(@clrizr-input), @@clrizr-primary-2-saturation, @@clrizr-primary-2-lightness);@clrizr-primary-1: hsl(hue(@clrizr-input), @@clrizr-primary-1-saturation, @@clrizr-primary-1-lightness);@clrizr-primary: @clrizr-input;@clrizr-complementary-1: hsl(@@clrizr-complementary-hue, @@clrizr-complementary-1-saturation, @@clrizr-complementary-1-lightness);@clrizr-complementary-2: hsl(@@clrizr-complementary-hue, saturation(@clrizr-input), lightness(@clrizr-input)); That’s is it, for the most part.  These variables now hold the theme for the one input color – @clrizr-input.  However, I have one last addition… Perceptive Luminance Well, after I got the colors, I decided I wanted to also get the best font color that would go on top of it.  Black or white depending on light or dark color.  Now I couldn’t just go with checking the lightness, as that is half the story.  You see, the human eye doesn’t see ALL colors equally well but rather has more cells for interpreting green light compared to blue or red.  So, using the ratio, we can calculate the perceptive luminance of each of the shades, and get the font color that best matches it! @clrizr-perceptive-luminance-ps2: round(1 - ( (0.299 * red(@clrizr-primary-2) ) + ( 0.587 * green(@clrizr-primary-2) ) + (0.114 * blue(@clrizr-primary-2)))/255)*255;@clrizr-perceptive-luminance-ps1: round(1 - ( (0.299 * red(@clrizr-primary-1) ) + ( 0.587 * green(@clrizr-primary-1) ) + (0.114 * blue(@clrizr-primary-1)))/255)*255;@clrizr-perceptive-luminance-ps: round(1 - ( (0.299 * red(@clrizr-primary) ) + ( 0.587 * green(@clrizr-primary) ) + (0.114 * blue(@clrizr-primary)))/255)*255;@clrizr-perceptive-luminance-pc1: round(1 - ( (0.299 * red(@clrizr-complementary-1)) + ( 0.587 * green(@clrizr-complementary-1)) + (0.114 * blue(@clrizr-complementary-1)))/255)*255;@clrizr-perceptive-luminance-pc2: round(1 - ( (0.299 * red(@clrizr-complementary-2)) + ( 0.587 * green(@clrizr-complementary-2)) + (0.114 * blue(@clrizr-complementary-2)))/255)*255; @clrizr-col-font-on-primary-2: rgb(@clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2);@clrizr-col-font-on-primary-1: rgb(@clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1);@clrizr-col-font-on-primary: rgb(@clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps);@clrizr-col-font-on-complementary-1: rgb(@clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1);@clrizr-col-font-on-complementary-2: rgb(@clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2); Conclusion That’s it!  I have posted a project on clrizr.codePlex.com for this, and included a testing page for you to test out how it works.  Feel free to use it in your own project, and if you have any questions, comments or suggestions, please feel free to leave them here as a comment, or on the contact page!

    Read the article

  • First PC Build (Part 1)

    - by Anthony Trudeau
    Originally posted on: http://geekswithblogs.net/tonyt/archive/2014/08/05/157959.aspxA couple of months ago I made the decision to build myself a new computer. The intended use is gaming and for using the last real version of Photoshop. I was motivated by the poor state of console gaming and a simple desire to do something I haven’t done before – build a PC from the ground up. I’ve been using PCs for more than two decades. I’ve replaced a component hear and there, but for the last 10 years or so I’ve only used laptops. Therefore, this article will be written from the perspective of someone familiar with PCs, but completely new at building. I’m not an expert and this is not a definitive guide for building a PC, but I do hope that it encourages you to try it yourself. Component List Research There was a lot of research necessary, because building a PC is completely new to me, and I haven’t kept up with what’s out there. The first thing you want to do is nail down what your goals are. Your goals are going to be driven by what you want to do with your computer and personal choice. Don’t neglect the second one, because if you’re doing this for fun you want to get what you want. In my case, I focused on three things: performance, longevity, and aesthetics. The performance aspect is important for gaming and Photoshop. This will drive what components you get. For example, heavy gaming use is going to drive your choice of graphics card. Longevity is relevant to me, because I don’t want to be changing things out anytime soon for the next hot game. The consequence of performance and longevity is cost. Finally, aesthetics was my next consideration. I could have just built a box, but it wouldn’t have been nearly as fun for me. Aesthetics might not be important to you. They are for me. I also like gadgets and that played into at least one purchase for this build. I used PC Part Picker to put together my component list. I found it invaluable during the process and I’d recommend it to everyone. One caveat is that I wouldn’t trust the compatibility aspects. It does a pretty good job of not steering you wrong, but do your own research. The rest of it isn’t really sexy. I started out with what appealed to me and then I made changes and additions as I dived deep into researching each component and interaction I could find. The resources I used are innumerable. I used reviews, product descriptions, forum posts (praises and problems), et al. to assist me. I also asked friends into gaming what they thought about my component list. And when I got near the end I posted my list to the Reddit /r/buildapc forum. I cannot stress the value of extra sets of eyeballs and first hand experiences. Some of the resources I used: PC Part Picker Tom’s Hardware bit-tech Reddit Purchase PC Part Picker favors certain vendors. You should look at others too. In my case I found their favorites to be the best. My priorities were out-the-door price and shipping time. I knew that once I started getting parts I’d want to start building. Luckily, I timed it well and everything arrived within the span of a few days. Here are my opinions on the vendors I ended up using in alphabetical order. Amazon.com is a good, reliable choice. They have excellent customer service in my experience, and I knew I wouldn’t have trouble with them. However, shipping time is often a problem when you use their free shipping unless you order expensive items (I’ve found items over $100 ship quickly). Ultimately though, price wasn’t always the best and their collection of sales tax in my state turned me off them. I did purchase my case from them. I ordered the mouse as well, but I cancelled after it was stuck four days in a “shipping soon” state. I purchased the mouse locally. Best Buy is not my favorite place to do business. There’s a lot of history with poor, uninterested sales representatives and they used to have a lot of bad anti-consumer policies. That’s a lot better now, but the bad taste is still in my mouth. I ended up purchasing the accessories from them including mouse (locally) and headphones. NCIX is a company that I’ve never heard of before. It popped up as a recommendation for my CPU cooler on PC Part Picker. I didn’t do a lot of research on the company, because their policy on you buying insurance for your orders turned me off. That policy makes it clear to me that the company finds me responsible for the shipment once it leaves their dock. That’s not right, and may run afoul of state laws. Regardless they shipped my CPU cooler quickly and I didn’t have a problem. NewEgg.com is a well known company. I had never done business with them, but I’m glad I did. They shipped quickly and provided good visibility over everything. The prices were also the best in most cases. My main complaint is that they have a lot of exchange only return policies on components. To their credit those policies are listed in the cart underneath each item. The visibility tells me that they’re not playing any shenanigans and made me comfortable dealing with that risk. The vast majority of what I ordered came from them. Coming Next In the next part I’ll tackle my build experience.

    Read the article

  • Building an OpenStack Cloud for Solaris Engineering, Part 1

    - by Dave Miner
    One of the signature features of the recently-released Solaris 11.2 is the OpenStack cloud computing platform.  Over on the Solaris OpenStack blog the development team is publishing lots of details about our version of OpenStack Havana as well as some tips on specific features, and I highly recommend reading those to get a feel for how we've leveraged Solaris's features to build a top-notch cloud platform.  In this and some subsequent posts I'm going to look at it from a different perspective, which is that of the enterprise administrator deploying an OpenStack cloud.  But this won't be just a theoretical perspective: I've spent the past several months putting together a deployment of OpenStack for use by the Solaris engineering organization, and now that it's in production we'll share how we built it and what we've learned so far.In the Solaris engineering organization we've long had dedicated lab systems dispersed among our various sites and a home-grown reservation tool for developers to reserve those systems; various teams also have private systems for specific testing purposes.  But as a developer, it can still be difficult to find systems you need, especially since most Solaris changes require testing on both SPARC and x86 systems before they can be integrated.  We've added virtual resources over the years as well in the form of LDOMs and zones (both traditional non-global zones and the new kernel zones).  Fundamentally, though, these were all still deployed in the same model: our overworked lab administrators set up pre-configured resources and we then reserve them.  Sounds like pretty much every traditional IT shop, right?  Which means that there's a lot of opportunity for efficiencies from greater use of virtualization and the self-service style of cloud computing.  As we were well into development of OpenStack on Solaris, I was recruited to figure out how we could deploy it to both provide more (and more efficient) development and test resources for the organization as well as a test environment for Solaris OpenStack.At this point, let's acknowledge one fact: deploying OpenStack is hard.  It's a very complex piece of software that makes use of sophisticated networking features and runs as a ton of service daemons with myriad configuration files.  The web UI, Horizon, doesn't often do a good job of providing detailed errors.  Even the command-line clients are not as transparent as you'd like, though at least you can turn on verbose and debug messaging and often get some clues as to what to look for, though it helps if you're good at reading JSON structure dumps.  I'd already learned all of this in doing a single-system Grizzly-on-Linux deployment for the development team to reference when they were getting started so I at least came to this job with some appreciation for what I was taking on.  The good news is that both we and the community have done a lot to make deployment much easier in the last year; probably the easiest approach is to download the OpenStack Unified Archive from OTN to get your hands on a single-system demonstration environment.  I highly recommend getting started with something like it to get some understanding of OpenStack before you embark on a more complex deployment.  For some situations, it may in fact be all you ever need.  If so, you don't need to read the rest of this series of posts!In the Solaris engineering case, we need a lot more horsepower than a single-system cloud can provide.  We need to support both SPARC and x86 VM's, and we have hundreds of developers so we want to be able to scale to support thousands of VM's, though we're going to build to that scale over time, not immediately.  We also want to be able to test both Solaris 11 updates and a release such as Solaris 12 that's under development so that we can work out any upgrade issues before release.  One thing we don't have is a requirement for extremely high availability, at least at this point.  We surely don't want a lot of down time, but we can tolerate scheduled outages and brief (as in an hour or so) unscheduled ones.  Thus I didn't need to spend effort on trying to get high availability everywhere.The diagram below shows our initial deployment design.  We're using six systems, most of which are x86 because we had more of those immediately available.  All of those systems reside on a management VLAN and are connected with a two-way link aggregation of 1 Gb links (we don't yet have 10 Gb switching infrastructure in place, but we'll get there).  A separate VLAN provides "public" (as in connected to the rest of Oracle's internal network) addresses, while we use VxLANs for the tenant networks. One system is more or less the control node, providing the MySQL database, RabbitMQ, Keystone, and the Nova API and scheduler as well as the Horizon console.  We're curious how this will perform and I anticipate eventually splitting at least the database off to another node to help simplify upgrades, but at our present scale this works.I had a couple of systems with lots of disk space, one of which was already configured as the Automated Installation server for the lab, so it's just providing the Glance image repository for OpenStack.  The other node with lots of disks provides Cinder block storage service; we also have a ZFS Storage Appliance that will help back-end Cinder in the near future, I just haven't had time to get it configured in yet.There's a separate system for Neutron, which is our Elastic Virtual Switch controller and handles the routing and NAT for the guests.  We don't have any need for firewalling in this deployment so we're not doing so.  We presently have only two tenants defined, one for the Solaris organization that's funding this cloud, and a separate tenant for other Oracle organizations that would like to try out OpenStack on Solaris.  Each tenant has one VxLAN defined initially, but we can of course add more.  Right now we have just a single /24 network for the floating IP's, once we get demand up to where we need more then we'll add them.Finally, we have started with just two compute nodes; one is an x86 system, the other is an LDOM on a SPARC T5-2.  We'll be adding more when demand reaches the level where we need them, but as we're still ramping up the user base it's less work to manage fewer nodes until then.My next post will delve into the details of building this OpenStack cloud's infrastructure, including how we're using various Solaris features such as Automated Installation, IPS packaging, SMF, and Puppet to deploy and manage the nodes.  After that we'll get into the specifics of configuring and running OpenStack itself.

    Read the article

  • Recursion in the form of a Recursive Func&lt;T, T&gt;

    - by ToStringTheory
    I gotta admit, I am kind of surprised that I didn’t realize I could do this sooner.  I recently had a problem which required a recursive function call to come up with the answer.  After some time messing around with a recursive method, and creating an API that I was not happy with, I was able to create an API that I enjoy, and seems intuitive. Introduction To bring it to a simple example, consider the summation to n: A mathematically identical formula is: In a .NET function, this can be represented by a function: Func<int, int> summation = x => x*(x+1)/2 Calling summation with an input integer will yield the summation to that number: var sum10 = summation(4); //sum10 would be equal to 10 But what if I wanted to get a second level summation…  First some to n, and then use that argument as the input to the same function, to find the second level summation: So as an easy example, calculate the summation to 3, which yields 6.  Then calculate the summation to 6 which yields 21. Represented as a mathematical formula - So what if I wanted to represent this as .NET functions.  I can always do: //using the summation formula from above var sum3 = summation(3); //sets sum3 to 6 var sum3_2 = summation(sum3); //sets sum3 to 21 I could always create a while loop to perform the calculations too: Func<int, int> summation = x => x*(x+1)/2; //for the interests of a smaller example, using shorthand int sumResultTo = 3; int level = 2; while(level-- > 0) { sumResultTo = summation(sumResultTo); } //sumResultTo is equal to 21 now. Or express it as a for-loop, method calls, etc…  I really didn’t like any of the options that I tried.  Then it dawned on me – since I was using a Func<T, T> anyways, why not use the Func’s output from one call as the input as another directly. Some Code So, I decided that I wanted a recursion class.  Something that I would be generic and reusable in case I ever wanted to do something like this again. It is limited to only the Func<T1, T2> level of Func, and T1 must be the same as T2. The first thing in this class is a private field for the function: private readonly Func<T, T> _functionToRecurse; So, I since I want the function to be unchangeable, I have defined it as readonly.  Therefore my constructor looks like: public Recursion(Func<T, T> functionToRecurse) { if (functionToRecurse == null) { throw new ArgumentNullException("functionToRecurse", "The function to recurse can not be null"); } _functionToRecurse = functionToRecurse; } Simple enough.  If you have any questions, feel free to post them in the comments, and I will be sure to answer them. Next, I want enough. If be able to get the result of a function dependent on how many levels of recursion: private Func<T, T> GetXLevel(int level) { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } if (level == 1) return _functionToRecurse; return _GetXLevel(level - 1, _functionToRecurse); } So, if you pass in 1 for the level, you get just the Func<T,T> back.  If you say that you want to go deeper down the rabbit hole, it calls a method which accepts the level it is at, and the function which it needs to use to recurse further: private Func<T, T> _GetXLevel(int level, Func<T, T> prevFunc) { if (level == 1) return y => prevFunc(_functionToRecurse(y)); return _GetXLevel(level - 1, y => prevFunc(_functionToRecurse(y))); } That is really all that is needed for this class. If I exposed the GetXLevel function publicly, I could use that to get the function for a level, and pass in the argument..  But I wanted something better.  So, I used the ‘this’ array operator for the class: public Func<T,T> this[int level] { get { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } return this.GetXLevel(level); } } So, using the same example above of finding the second recursion of the summation of 3: var summator = new Recursion<int>(x => (x * (x + 1)) / 2); var sum_3_level2 = summator[2](3); //yields 21 You can even find just store the delegate to the second level summation, and use it multiple times: var summator = new Recursion<int>(x => (x * (x + 1)) / 2); var sum_level2 = summator[2]; var sum_3_level2 = sum_level2(3); //yields 21 var sum_4_level2 = sum_level2(4); //yields 55 var sum_5_level2 = sum_level2(5); //yields 120 Full Code Don’t think I was just going to hold off on the full file together and make you do the hard work…  Copy this into a new class file: public class Recursion<T> { private readonly Func<T, T> _functionToRecurse; public Recursion(Func<T, T> functionToRecurse) { if (functionToRecurse == null) { throw new ArgumentNullException("functionToRecurse", "The function to recurse can not be null"); } _functionToRecurse = functionToRecurse; } public Func<T,T> this[int level] { get { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } return this.GetXLevel(level); } } private Func<T, T> GetXLevel(int level) { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } if (level == 1) return _functionToRecurse; return _GetXLevel(level - 1, _functionToRecurse); } private Func<T, T> _GetXLevel(int level, Func<T, T> prevFunc) { if (level == 1) return y => prevFunc(_functionToRecurse(y)); return _GetXLevel(level - 1, y => prevFunc(_functionToRecurse(y))); } } Conclusion The great thing about this class, is that it can be used with any function with same input/output parameters.  I strived to find an implementation that I found clean and useful, and I finally settled on this.  If you have feedback – good or bad, I would love to hear it!

    Read the article

  • Feedback on meeting of the Linux User Group of Mauritius

    Once upon a time in a country far far away... Okay, actually it's not that bad but it has been a while since the last meeting of the Linux User Group of Mauritius (LUGM). There have been plans in the past but it never really happened. Finally, Selven took the opportunity and organised a new meetup with low administrative overhead, proper scheduling on alternative dates and a small attendee's survey on the preferred option. All the pre-work was nicely executed. First, I wasn't sure whether it would be possible to attend. Luckily I got some additional information, like children should come, too, and I was sold to this community gathering. According to other long-term members of the LUGM it was the first time 'ever' that a gathering was organised outside of Quatre Bornes, and I have to admit it was great! LUGM - user group meeting on the 15.06.2013 in L'Escalier Quick overview of Linux & the LUGM With a little bit of delay the LUGM meeting officially started with a quick overview and introduction to Linux presented by Avinash. During the session he told the audience that there had been quite some activity over the island some years ago but unfortunately it had been quiet during recent times. Of course, we also spoke about the acknowledged world dominance of Linux - thanks to Android - and the interesting possibilities for countries like Mauritius. It is known that a couple of public institutions have there back-end infrastructure running on Red Hat Linux systems but the presence on the desktop is still very low. Users are simply hanging on to Windows XP and older versions of Microsoft Office. Following the introduction of the LUGM Ajay joined into the session and it quickly changed into a panel discussion with lots of interesting questions and answers, sharing of first-hand experience either on the job or in private use of Linux, and a couple of ideas about how the LUGM could promote Linux a bit more in Mauritius. It was great to get an insight into other attendee's opinion and activities. Especially taking into consideration that I'm already using Linux since around 1996/97. Frankly speaking, I bought a SuSE 4.x distribution back in those days because I couldn't achieve certain tasks on Windows NT 4.0 without spending a fortune. OpenELEC Mediacenter Next, Selven gave us decent introduction on OpenELEC: Open Embedded Linux Entertainment Center (OpenELEC) is a small Linux distribution built from scratch as a platform to turn your computer into an XBMC media center. OpenELEC is designed to make your system boot fast, and the install is so easy that anyone can turn a blank PC into a media machine in less than 15 minutes. I didn't know about it until this presentation. In the past, I was mainly attached to Video Disk Recorder (VDR) as it allows the use of satellite receiver cards very easily. Hm, somehow I'm still missing my precious HTPC that I had to leave back in Germany years ago. It was great piece of hardware and software; self-built PC in a standard HiFi-sized (43cm) black desktop casing with 2 full-featured Hauppauge DVB-s cards, an old-fashioned Voodoo graphics card, WiFi card, Pioneer slot-in DVD drive, and fully remote controlled via infra-red thanks to Debian, VDR and LIRC. With EP Guide, scheduled recordings and general multimedia centre it offered all the necessary comfort in the living room, besides a Nintendo game console; actually a GameCube at that time... But I have to admit that putting OpenELEC on a Raspberry Pi would be a cool DIY project in the near future. LUGM - our next generation of linux users (15.06.2013) Project Evil Genius (PEG) Don't be scared of the paragraph header. Ish gave us a cool explanation why he named it PEG - Project Evil Genius; it's because of the time of the day when he was scripting down his ideas to be able to build, package and provide software applications to various Linux distributions. The main influence came from openSuSE but the platform didn't cater for his needs and ideas, so he started to work out something on his own. During his passionate session he also talked about the amazing experience he had due to other Linux users from all over the world. During the next couple of days Ish promised to put his script to GitHub... Looking forward to that. Check out Ish's personal blog over at hacklog.in. Highly recommended to read. Why India? Simply because the registration fees per year for an Indian domain are approximately 20 times less than for a Mauritian domain (.mu). Exploring the beach of L'Escalier af the meeting 'After-party' at the beach of L'Escalier Puh, after such interesting sessions, ideas around Linux and good conversation during the breaks and over lunch it was time for a little break-out. Selven suggested that we all should head down to the beach of L'Escalier and get some impressions of nature down here in the south of the island. Talking about 'beach' ;-) - absolutely not comparable to the white-sanded ones here in Flic en Flac... There are no lagoons down at the south coast of Mauriitus, and watching the breaking waves is a different experience and joy after all. Unfortunately, I was a little bit worried about the thoughtless littering at such a remote location. You have to drive on natural paths through the sugar cane fields and I was really shocked by the amount of rubbish lying around almost everywhere. Sad, really sad and it concurs with Yasir's recent article on the same topic. Resumé & outlook It was a great event. I met with new people, had some good conversations, and even my children enjoyed themselves the whole day. The location was well-chosen, enough space for each and everyone, parking spaces and even a playground for the children. Also, a big "Thank You" to Selven and his helpers for the organisation and preparation of lunch. I'm kind of sure that this was an exceptional meeting of LUGM and I'm really looking forward to the next gathering of Linux geeks. Hopefully, soon. All images are courtesy of Avinash Meetoo. More pictures are available on Flickr.

    Read the article

  • parallel_for_each from amp.h – part 1

    - by Daniel Moth
    This posts assumes that you've read my other C++ AMP posts on index<N> and extent<N>, as well as about the restrict modifier. It also assumes you are familiar with C++ lambdas (if not, follow my links to C++ documentation). Basic structure and parameters Now we are ready for part 1 of the description of the new overload for the concurrency::parallel_for_each function. The basic new parallel_for_each method signature returns void and accepts two parameters: a grid<N> (think of it as an alias to extent) a restrict(direct3d) lambda, whose signature is such that it returns void and accepts an index of the same rank as the grid So it looks something like this (with generous returns for more palatable formatting) assuming we are dealing with a 2-dimensional space: // some_code_A parallel_for_each( g, // g is of type grid<2> [ ](index<2> idx) restrict(direct3d) { // kernel code } ); // some_code_B The parallel_for_each will execute the body of the lambda (which must have the restrict modifier), on the GPU. We also call the lambda body the "kernel". The kernel will be executed multiple times, once per scheduled GPU thread. The only difference in each execution is the value of the index object (aka as the GPU thread ID in this context) that gets passed to your kernel code. The number of GPU threads (and the values of each index) is determined by the grid object you pass, as described next. You know that grid is simply a wrapper on extent. In this context, one way to think about it is that the extent generates a number of index objects. So for the example above, if your grid was setup by some_code_A as follows: extent<2> e(2,3); grid<2> g(e); ...then given that: e.size()==6, e[0]==2, and e[1]=3 ...the six index<2> objects it generates (and hence the values that your lambda would receive) are:    (0,0) (1,0) (0,1) (1,1) (0,2) (1,2) So what the above means is that the lambda body with the algorithm that you wrote will get executed 6 times and the index<2> object you receive each time will have one of the values just listed above (of course, each one will only appear once, the order is indeterminate, and they are likely to call your code at the same exact time). Obviously, in real GPU programming, you'd typically be scheduling thousands if not millions of threads, not just 6. If you've been following along you should be thinking: "that is all fine and makes sense, but what can I do in the kernel since I passed nothing else meaningful to it, and it is not returning any values out to me?" Passing data in and out It is a good question, and in data parallel algorithms indeed you typically want to pass some data in, perform some operation, and then typically return some results out. The way you pass data into the kernel, is by capturing variables in the lambda (again, if you are not familiar with them, follow the links about C++ lambdas), and the way you use data after the kernel is done executing is simply by using those same variables. In the example above, the lambda was written in a fairly useless way with an empty capture list: [ ](index<2> idx) restrict(direct3d), where the empty square brackets means that no variables were captured. If instead I write it like this [&](index<2> idx) restrict(direct3d), then all variables in the some_code_A region are made available to the lambda by reference, but as soon as I try to use any of those variables in the lambda, I will receive a compiler error. This has to do with one of the direct3d restrictions, where only one type can be capture by reference: objects of the new concurrency::array class that I'll introduce in the next post (suffice for now to think of it as a container of data). If I write the lambda line like this [=](index<2> idx) restrict(direct3d), all variables in the some_code_A region are made available to the lambda by value. This works for some types (e.g. an integer), but not for all, as per the restrictions for direct3d. In particular, no useful data classes work except for one new type we introduce with C++ AMP: objects of the new concurrency::array_view class, that I'll introduce in the post after next. Also note that if you capture some variable by value, you could use it as input to your algorithm, but you wouldn’t be able to observe changes to it after the parallel_for_each call (e.g. in some_code_B region since it was passed by value) – the exception to this rule is the array_view since (as we'll see in a future post) it is a wrapper for data, not a container. Finally, for completeness, you can write your lambda, e.g. like this [av, &ar](index<2> idx) restrict(direct3d) where av is a variable of type array_view and ar is a variable of type array - the point being you can be very specific about what variables you capture and how. So it looks like from a large data perspective you can only capture array and array_view objects in the lambda (that is how you pass data to your kernel) and then use the many threads that call your code (each with a unique index) to perform some operation. You can also capture some limited types by value, as input only. When the last thread completes execution of your lambda, the data in the array_view or array are ready to be used in the some_code_B region. We'll talk more about all this in future posts… (a)synchronous Please note that the parallel_for_each executes as if synchronous to the calling code, but in reality, it is asynchronous. I.e. once the parallel_for_each call is made and the kernel has been passed to the runtime, the some_code_B region continues to execute immediately by the CPU thread, while in parallel the kernel is executed by the GPU threads. However, if you try to access the (array or array_view) data that you captured in the lambda in the some_code_B region, your code will block until the results become available. Hence the correct statement: the parallel_for_each is as-if synchronous in terms of visible side-effects, but asynchronous in reality.   That's all for now, we'll revisit the parallel_for_each description, once we introduce properly array and array_view – coming next. Comments about this post by Daniel Moth welcome at the original blog.

    Read the article

  • Notifications for Expiring DBSNMP Passwords

    - by Courtney Llamas
    Most user accounts these days have a password profile on them that automatically expires the password after a set number of days.   Depending on your company’s security requirements, this may be as little as 30 days or as long as 365 days, although typically it falls between 60-90 days. For a normal user, this can cause a small interruption in your day as you have to go get your password reset by an admin. When this happens to privileged accounts, such as the DBSNMP account that is responsible for monitoring database availability, it can cause bigger problems. In Oracle Enterprise Manager 12c you may notice the error message “ORA-28002: the password will expire within 5 days” when you connect to a target, or worse you may get “ORA-28001: the password has expired". If you wait too long, your monitoring will fail because the password is locked out. Wouldn’t it be nice if we could get an alert 10 days before our DBSNMP password expired? Thanks to Oracle Enterprise Manager 12c Metric Extensions (ME), you can! See the Oracle Enterprise Manager Cloud Control Administrator’s Guide for more information on Metric Extensions. To create a metric extension, select Enterprise / Monitoring / Metric Extensions, and then click on Create. On the General Properties screen select either Cluster Database or Database Instance, depending on which target you need to monitor.  If you have both RAC and Single instance you may need to create one for each. In this example we will create a Cluster Database metric.  Enter a Name for the ME and a Display Name. Then select SQL for the Adapter.  Adjust the Collection Schedule as desired, for this example we will collect this metric every 1 day. Notice for metric collected every day, we can determine the exact time we want to collect. On the Adapter page, enter the query that you wish to execute.  In this example we will use the query below that specifically checks for the DBSNMP user that is expiring within 10 days. Of course, you can adjust this query to alert for any user that can cause an outage such as an application account or service account such as RMAN. select username, account_status, trunc(expiry_date-sysdate) days_to_expirefrom dba_userswhere username = 'DBSNMP'and expiry_date is not null; The next step is to create the columns to store the data returned from the query.  Click Add and add a column for each of the fields in the same order that data is returned.  The table below will help you complete the column additions. Name Display Name Column Type Value Type Metric Category Unit Username User Name Key String Security AccountStatus Account Status Data String Security DaysToExpire Days Until Expiration Data Number Security Days When creating the DaysToExpire column, you can add a default threshold here for Warning and Critical (say < 10 and 5).  When all columns have been added, click Next. On the Credentials page, you can choose to use the default monitoring credentials or specify new credentials.  We will use the default credentials established for our target (dbsnmp). The next step is to test your Metric Extension.  Click on Add to select a target for testing, then click Select. Now click the button Run Test to execute the test against the selected target(s). We can see in the example below that the Metric Extension has executed and returned a value of 68 days to expire. Click Next to proceed. Review the metric extension in the final screen and click Finish. The metric will be created in Editable status.  Select the metric, click Actions and select Deployable Draft. You can do this once more to move to Published. Finally, we want to apply this metric to a target. When managing many targets, it’s best to add your metric to a template, for details on adding a Metric Extension to a template see the Administrator’s Guide. For this example, we will deploy this to a target directly. Select Actions / Deploy to Targets. Click Add and select the target you wish to deploy to and click Submit.  Once deployment is complete, we can go to the target and view the Metric & Collection Settings to see the new metric and its thresholds.   After some time, you will find the metric has collected and the days to expiration for DBSNMP user can be seen in the All Metrics view.   For metrics collected once per day, you may have to wait up to 24 hours to see the metric and current severity. In the example below, the current severity is Clear (green check) as it is not scheduled to expire within 10 days. To test the notification, we can edit the thresholds for the new metric so they trigger an alert.  Our password expires in 139 days, so we’ll change our Warning to 140 and leave Critical at 5, in our example we also changed the collection time to every 5 minutes.  At the next collection, you’ll find that the current severity changes to a Warning and any related Incident Rules would be triggered to create an Incident or Notification as desired. Now that you get a notification that your DBSNMP passwords is about to expire, you can use OEM Command Line Interface (EM CLI) verb update_db_password to change it at both the database target and the OEM target in one step.  The caveat is you must know the existing password to use the update_db_password command.  To learn more about EM CLI, see the Oracle Enterprise Manager Command Line Interface Guide.  Below is an example of changing the password with the update_db_password verb.  $ ./emcli update_db_password -target_name=emrep -target_type=oracle_database -user_name=dbsnmp -change_at_target=yes -change_all_references=yes Enter value for old_password :Enter value for new_password :Enter value for retype_new_password :Successfully submitted a job to change the password in Enterprise Manager and on the target database: "emrep"Execute "emcli get_jobs -job_id=FA66C1C4D663297FE0437656F20ACC84" to check the status of the job.Search for job name "CHANGE_PWD_JOB_FA66C1C4D662297FE0437656F20ACC84" on the Jobs home page to check job execution details. The subsequent job created will typically run quickly enough that a blackout is not needed, however if you submit a script with many targets to change, your job may run slower so adding a blackout to the script is recommended. $ ./emcli get_jobs -job_id=FA66C1C4D663297FE0437656F20ACC84 Name Type Job ID Execution ID Scheduled Completed TZ Offset Status Status ID Owner Target Type Target Name CHANGE_PWD_JOB_FA66C1C4D662297FE0437656F20ACC84 ChangePassword FA66C1C4D663297FE0437656F20ACC84 FA66C1C4D665297FE0437656F20ACC84 2014-05-28 09:39:12 2014-05-28 09:39:18 GMT-07:00 Succeeded 5 SYSMAN oracle_database emrep After implementing the above Metric Extension and using the EM CLI update_db_password verb, you will be able to stay on top of your DBSNMP password changes without experiencing an unplanned monitoring outage.  

    Read the article

  • The code works but when using printf it gives me a weird answer. Help please [closed]

    - by user71458
    //Programmer-William Chen //Seventh Period Computer Science II //Problem Statement - First get the elapsed times and the program will find the //split times for the user to see. // //Algorithm- First the programmer makes the prototype and calls them in the //main function. The programmer then asks the user to input lap time data. //Secondly, you convert the splits into seconds and subtract them so you can //find the splits. Then the average is all the lap time's in seconds. Finally, //the programmer printf all the results for the user to see. #include <iostream> #include <stdlib.h> #include <math.h> #include <conio.h> #include <stdio.h> using namespace std; void thisgetsElapsedTimes( int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5); //this is prototype void thisconvertstoseconds ( int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5, int &split1, int &split2, int &split3, int &split4, int &split5);//this too void thisfindsSplits(int &m1, int &m2, int &m3, int &m4, int &m5, int &split1, int &split2, int &split3, int &split4, int &split5, int &split6, int &split7, int &split8, int &split9, int &split10);// this is part of prototype void thisisthesecondconversation (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S,int &split2S, int &split3S, int &split4S, int &split5S, int &split1, int &split2, int &split3, int &split4, int &split5);//this gets a value void thisfindstheaverage(double &average, int &split1, int &split2, int &split3, int &split4, int &split5);//and this void thisprintsstuff( int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S, int &split2S, int &split3S, int &split4S, int &split5S, double &average); //this prints int main(int argc, char *argv[]) { int m1, m2, m3, m4, m5, s1, s2, s3, s4, s5, split1, split2, split3, split4, split5, split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S; int split6, split7, split8, split9, split10; double average; char thistakescolon; thisgetsElapsedTimes ( m1, m2, m3, m4, m5, s1, s2, s3, s4, s5); thisconvertstoseconds ( m1, m2, m3, m4, m5, s1, s2, s3, s4, s5, split1, split2, split3, split4, split5); thisfindsSplits ( m1, m2, m3, m4, m5, split1, split2, split3, split4, split5, split6, split7, split8, split9, split10); thisisthesecondconversation ( split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S, split1, split2, split3, split4, split5); thisfindstheaverage ( average, split1, split2, split3, split4, split5); thisprintsstuff ( split1M, split2M, split3M, split4M, split5M, split1S, split2S, split3S, split4S, split5S, average); // these are calling statements and they call from the main function to the other functions. system("PAUSE"); return 0; } void thisgetsElapsedTimes(int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5) { char thistakescolon; cout << "Enter the elapsed time:" << endl; cout << " Kilometer 1 "; cin m1 thistakescolon s1; cout << " Kilometer 2 "; cin m2 thistakescolon s2; cout << " Kilometer 3 " ; cin m3 thistakescolon s3; cout << " Kilometer 4 "; cin m4 thistakescolon s4; cout << " Kilometer 5 "; cin m5 thistakescolon s5; // this gets the data required to get the results needed for the user to see // . } void thisconvertstoseconds (int &m1, int &m2, int &m3, int &m4, int &m5, int &s1, int &s2, int &s3, int &s4, int &s5, int &split1, int &split2, int &split3, int &split4, int &split5) { split1 = (m1 * 60) + s1;//this converts for minutes to seconds for m1 split2 = (m2 * 60) + s2;//this converts for minutes to seconds for m2 split3 = (m3 * 60) + s3;//this converts for minutes to seconds for m3 split4 = (m4 * 60) + s4;//this converts for minutes to seconds for m4 split5 = (m5 * 60) + s5;//this converts for minutes to seconds for m5 } void thisfindsSplits (int &m1, int &m2, int &m3, int &m4, int &m5,int &split1, int &split2, int &split3, int &split4, int &split5, int &split6, int &split7, int &split8, int &split9, int &split10)//this is function heading { split6 = split1; //this is split for the first lap. split7 = split2 - split1;//this is split for the second lap. split8 = split3 - split2;//this is split for the third lap. split9 = split4 - split3;//this is split for the fourth lap. split10 = split5 - split4;//this is split for the fifth lap. } void thisfindstheaverage(double &average, int &split1, int &split2, int &split3, int &split4, int &split5) { average = (split1 + split2 + split3 + split4 + split5)/5; // this finds the average from all the splits in seconds } void thisisthesecondconversation (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S,int &split2S, int &split3S, int &split4S, int &split5S, int &split1, int &split2, int &split3, int &split4, int &split5) { split1M = split1 * 60; //this finds the split times split1S = split1M - split1 * 60; //then this finds split2M = split2 * 60; //and all of this split2S = split2M - split2 * 60; //does basically split3M = split3 * 60; //the same thing split3S = split3M - split3 * 60; //all of it split4M = split4 * 60; //it's also a split4S = split4M - split4 * 60; //function split5M = split5 * 60; //and it finds the splits split5S = split5M - split5 * 60; //for each lap. } void thisprintsstuff (int &split1M, int &split2M, int &split3M, int &split4M, int &split5M, int &split1S, int &split2S, int &split3S, int &split4S, int &split5S, double &average)// this is function heading { printf("\n kilometer 1 %d" , ":02%d",'split1M','split1S'); printf("\n kilometer 2 %d" , ":02%d",'split2M','split2S'); printf("\n kilometer 3 %d" , ":02%d",'split3M','split3S'); printf("\n kilometer 4 %d" , ":02%d",'split4M','split4S'); printf("\n kilometer 5 %d" , ":02%d",'split5M','split5S'); printf("\n your average pace is ",'average',"per kilometer \n", "William Chen\n"); // this printf so the programmer // can allow the user to see // the results from the data gathered. }

    Read the article

  • Getting Started With Tailoring Business Processes

    - by Richard Bingham
    In this article, and for the sake of simplicity, we will use the term “On-Premise” to mean a deployment where you have design-time development access to the instance, including administration of the technology components, the applications filesystem, and the database. In reality this might be a local development instance that is then supported by a team who can deploy your customizations to the restricted production instance equivalents. Tools Overview Firstly let’s look at the Design-Time tools within JDeveloper for customizing and extending the artifacts of a Business Process. In essence this falls into two buckets; SOA Composite Editor for working with BPEL processes, and the BPM Studio. The SOA Composite Editor As a standard extension to JDeveloper, this graphical design tool should be familiar to anyone previously worked with Oracle SOA Server. With easy-to-use modeling capability, backed-up by full XML source-view (for read-only), it provides everything that is needed to implement the technical design. In simple terms, once deployed to the remote SOA Server the composite components (like Mediator) leverage the Event Delivery Network (EDN) for interaction with the application logic. If you are customizing an existing Fusion Applications BPEL process then be aware that it does support MDS-based customization layers just like Page Composer where different customizations are used based on the run-time context, like for a specific Product or Business Unit. This also makes them safe from patching and upgrades, although only a single active version of the composite is available at run-time. This is defined by a field on the composite record, available in Enterprise Manager. Obviously if you wish to fire different activities and tasks based on the user context then you can should include switches to fork the flows in your custom BPEL process. Figure 1 – A BPEL process in Composite Editor The following describes the simplified steps for making customizations to BPEL processes. This is the most common method of changing the business processes of Fusion Applications, as over 400 BPEL-based composite applications are provided out-of-the-box. Setup your local Fusion Applications JDeveloper environment. The SOA Composite Editor should be installed as part of the Fusion Applications extension. If there are problems you can also find it under the ‘Check for Updates’ help menu option. Since SOA Server is not part of the JDeveloper integrated WebLogic Server, setup a standalone WebLogic environment for deploying and testing. Obviously you might use a Fusion Applications development instance also. Package the existing standard Fusion Applications SOA Composite using Enterprise Manager and export it as a complete SOA Archive (SAR) file, resulting in a local .jar file. You may need to ask your system administrator for this. Import the exported SAR .jar file into JDeveloper using the File menu, under the option ‘SOA Archive into SOA Project’. In JDeveloper set the appropriate customization layer values, and then change from the default role to the Fusion Applications Customization Developer role. Make the customizations and save the application project. Finally redeploy the composite application, either to a direct Application Server connection, or as a fresh SAR (jar) file that can then be re-imported and deployed via Enterprise Manager. The Business Process Management (BPM) Suite In addition to the relatively low-level development environment associated with BPEL process creation, Oracle provides a suite of products that allow business process adjustments to be made without the need for some of the programming skills.  The aim is to abstract much of the technical implementation and to provide a Business Analyst tools for immediately implementing organization changes. Obviously there are some limitations on what they can do, however the BPM Suite functionality increases with each release and for the majority of the cases the tools remains as applicable as its developer-orientated sister. At the current time business processes must be explicitly coded to support just one of these use-cases, either BPEL for developer use or BPM for business analyst use. That said, they both run on the same SOA Server in much the same way. The components bundled in each SOA Composite Application can be verified by inspection through Enterprise Manager. Figure 2 – A BPM Process in JDeveloper BPM Suite. BPM processes are written in a standard notation (BPMN) and the modeling tools are very similar to that of BPEL. The steps to deploy a custom BPM process are also essentially much the same, since the BPM process is bundled into a SOA Composite just like a BPEL process. As such the SOA Composite Editor  actually has support for both artifacts and even allows use of them together, such as a calling a BPM process as a partnerlink from a BPEL process. For more details see the references below. Business Analyst Tooling In addition to using JDeveloper extensions for BPM development, there are run-time tools that Business Analysts can use to make adjustments, so that without high costs of an IT project the system can be tuned to match changes to the business operation. The first tool to consider is the BPM Composer, deployed with the middleware SOA Server and accessible online, and for Fusion Applications it is under the Business Process icon on the homepage of the Application Composer. Figure 3 – Business Process Composer showing a CRM process flow. The key difference between this and using JDeveloper is that the BPM Composer has a Business Catalog prepopulated with features and functions that can be used, mostly through registered WebServices. This means no coding or complex interface development is required, simply drag-drop-configure. The items in the business catalog are seeded by either Oracle (as a BPM Template) or added to by your own custom development. You cannot create or generate catalog content from BPM Composer directly. As per the screenshot you can see the Business Catalog content in the BPM Project browser region. In addition, other online tools for use by Business Analysts include the BPM Worklist application for editing business rules and approval management configuration, plus the SOA Composer which focuses on non-approval business rules and domain value maps. At the current time there are only a handful of BPM processes shipped with Fusion Applications HCM and CRM, including on-boarding workers and processing customer registrations.  This also means a limited number of associated BPM Templates provided out-of-the-box, therefore a limited Business Catalog. That said, BPM-based extension is a powerful capability to leverage and will most likely develop going forwards, especially for use in SaaS deployments where full design-time JDeveloper access is not available. Further Reading For BPEL – Fusion Applications Extensibility Guide – Section 12 For BPM – Fusion Applications Extensibility Guide – Section 7 The product-specific documentation and implementation guides for Fusion Applications Fusion Middleware Developers Guide for SOA Suite Modeling and Implementation Guide for Oracle Business Process Management User’s Guide for Oracle Business Process Composer Oracle University courses on BPM Suite and SOA Development

    Read the article

  • Resolving data redundancy up front

    - by okeofs
    Introduction As all of us do when confronted with a problem, the resource of choice is to ‘Google it’. This is where the plot thickens. Recently I was asked to stage data from numerous databases which were to be loaded into a data warehouse. To make a long story short, I was looking for a manner in which to obtain the table names from each database, to ascertain potential overlap.   As the source data comes from a SQL database created from dumps of a third party product,  one could say that there were +/- 95 tables for each database.   Yes I know that first instinct is to use the system stored procedure “exec sp_msforeachdb 'select "?" AS db, * from [?].sys.tables'”. However, if one stops to think about this, it would be nice to have all the results in a temporary or disc based  table; which in itself , implies additional labour. This said,  I decided to ‘re-invent’ the wheel. The full code sample may be found at the bottom of this article.   Define a few temporary tables and variables   declare @SQL varchar(max); declare @databasename varchar(75) /* drop table ##rawdata3 drop table #rawdata1 drop table #rawdata11 */ -- A temp table to hold the names of my databases CREATE TABLE #rawdata1 (    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) )     --A temp table with the same database names as above, HOWEVER using an --Identity number (recNO) as a loop variable. --You will note below that I loop through until I reach 25 (see below) as at --that point the system databases, the reporting server database etc begin. --1- 24 are user databases. These are really what I was looking for. --Whilst NOT the best solution,it works and the code was meant as a quick --and dirty. CREATE TABLE #rawdata11 (    recNo int identity(1,1),    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) )   --My output table showing the database name and table name CREATE TABLE ##rawdata3 (    database_name varchar(75) ,    table_name varchar(75), )   Insert the database names into a temporary table I pull the database names using the system stored procedure sp_databases   INSERT INTO #rawdata1 EXEC sp_databases Go   Insert the results from #rawdata1 into a table containing a record number  #rawdata11 so that I can LOOP through the extract   INSERT into #rawdata11 select * from  #rawdata1   We now declare 3 more variables:  @kounter is used to keep track of our position within the loop. @databasename is used to keep track of the’ current ‘ database name being used in the current pass of the loop;  as inorder to obtain the tables for that database we  need to issue a ‘USE’ statement, an insert command and other related code parts. This is the challenging part. @sql is a varchar(max) variable used to contain the ‘USE’ statement PLUS the’ insert ‘ code statements. We now initalize @kounter to 1 .   declare @kounter int; declare @databasename varchar(75); declare @sql varchar(max); set @kounter = 1   The Loop The astute reader will remember that the temporary table #rawdata11 contains our  database names  and each ‘database row’ has a record number (recNo). I am only interested in record numbers under 25. I now set the value of the temporary variable @DatabaseName (see below) .Note that I used the row number as a part of the predicate. Now, knowing the database name, I can create dynamic T-SQL to be executed using the sp_sqlexec stored procedure (see the code in red below). Finally, after all the tables for that given database have been placed in temporary table ##rawdata3, I increment the counter and continue on. Note that I used a global temporary table to ensure that the result set persists after the termination of the run. At some stage, I plan to redo this part of the code, as global temporary tables are not really an ideal solution.    WHILE (@kounter < 25)  BEGIN  select @DatabaseName = database_name from #rawdata11 where recNo = @kounter  set @SQL = 'Use ' + @DatabaseName + ' Insert into ##rawdata3 ' + + ' SELECT table_catalog,Table_name FROM information_schema.tables' exec sp_sqlexec  @Sql  SET @kounter  = @kounter + 1  END   The full code extract   Here is the full code sample.   declare @SQL varchar(max); declare @databasename varchar(75) /* drop table ##rawdata3 drop table #rawdata1 drop table #rawdata11 */ CREATE TABLE #rawdata1 (    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) ) CREATE TABLE #rawdata11 (    recNo int identity(1,1),    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) ) CREATE TABLE ##rawdata3 (    database_name varchar(75) ,    table_name varchar(75), )   INSERT INTO #rawdata1 EXEC sp_databases go INSERT into #rawdata11 select * from  #rawdata1 declare @kounter int; declare @databasename varchar(75); declare @sql varchar(max); set @kounter = 1 WHILE (@kounter < 25)  BEGIN  select @databasename = database_name from #rawdata11 where recNo = @kounter  set @SQL = 'Use ' + @DatabaseName + ' Insert into ##rawdata3 ' + + ' SELECT table_catalog,Table_name FROM information_schema.tables' exec sp_sqlexec  @Sql  SET @kounter  = @kounter + 1  END    select * from ##rawdata3  where table_name like '%SalesOrderHeader%'

    Read the article

  • Making Those PanelBoxes Behave

    - by Duncan Mills
    I have a little problem to solve earlier this week - misbehaving <af:panelBox> components... What do I mean by that? Well here's the scenario, I have a page fragment containing a set of panelBoxes arranged vertically. As it happens, they are stamped out in a loop but that does not really matter. What I want to be able to do is to provide the user with a simple UI to close and open all of the panelBoxes in concert. This could also apply to showDetailHeader and similar items with a disclosed attrubute, but in this case it's good old panelBoxes.  Ok, so the basic solution to this should be self evident. I can set up a suitable scoped managed bean that the panelBoxes all refer to for their disclosed attribute state. Then the open all / close commandButtons in the UI can simply set the state of that bean for all the panelBoxes to pick up via EL on their disclosed attribute. Sound OK? Well that works basically without a hitch, but turns out that there is a slight problem and this is where the framework is attempting to be a little too helpful. The issue is that is the user manually discloses or hides a panelBox then that will override the value that the EL is setting. So for example. I start the page with all panelBoxes collapsed, all set by the EL state I'm storing on the session I manually disclose panelBox no 1. I press the Expand All button - all works as you would hope and all the panelBoxes are now disclosed, including of course panelBox 1 which I just expanded manually. Finally I press the Collapse All button and everything collapses except that first panelBox that I manually disclosed.  The problem is that the component remembers this manual disclosure and that overrides the value provided by the expression. If I change the viewId (navigate away and back) then the panelBox will start to behave again, until of course I touch it again! Now, the more astute amoungst you would think (as I did) Ah, sound like the MDS personalizaton stuff is getting in the way and the solution should simply be to set the dontPersist attribute to disclosed | ALL. Alas this does not fix the issue.  After a little noodling on the best way to approach this I came up with a solution that works well, although if you think of an alternative way do let me know. The principle is simple. In the disclosureListener for the panelBox I take a note of the clientID of the panelBox component that has been touched by the user along with the state. This all gets stored in a Map of Booleans in ViewScope which is keyed by clientID and stores the current disclosed state in the Boolean value.  The listener looks like this (it's held in a request scope backing bean for the page): public void handlePBDisclosureEvent(DisclosureEvent disclosureEvent) { String clientId = disclosureEvent.getComponent().getClientId(FacesContext.getCurrentInstance()); boolean state = disclosureEvent.isExpanded(); pbState.addTouchedPanelBox(clientId, state); } The pbState variable referenced here is a reference to the bean which will hold the state of the panelBoxes that lives in viewScope (recall that everything is re-set when the viewid is changed so keeping this in viewScope is just fine and cleans things up automatically). The addTouchedPanelBox() method looks like this: public void addTouchedPanelBox(String clientId, boolean state) { //create the cache if needed this is just a Map<String,Boolean> if (_touchedPanelBoxState == null) { _touchedPanelBoxState = new HashMap<String, Boolean>(); } // Simply put / replace _touchedPanelBoxState.put(clientId, state); } So that's the first part, we now have a record of every panelBox that the user has touched. So what do we do when the Collapse All or Expand All buttons are pressed? Here we do some JavaScript magic. Basically for each clientID that we have stored away, we issue a client side disclosure event from JavaScript - just as if the user had gone back and changed it manually. So here's the Collapse All button action: public String CloseAllAction() { submitDiscloseOverride(pbState.getTouchedClientIds(true), false); _uiManager.closeAllBoxes(); return null; }  The _uiManager.closeAllBoxes() method is just manipulating the master-state that all of the panelBoxes are bound to using EL. The interesting bit though is the line:  submitDiscloseOverride(pbState.getTouchedClientIds(true), false); To break that down, the first part is a call to that viewScoped state holder to ask for a list of clientIDs that need to be "tweaked": public String getTouchedClientIds(boolean targetState) { StringBuilder sb = new StringBuilder(); if (_touchedPanelBoxState != null && _touchedPanelBoxState.size() > 0) { for (Map.Entry<String, Boolean> entry : _touchedPanelBoxState.entrySet()) { if (entry.getValue() == targetState) { if (sb.length() > 0) { sb.append(','); } sb.append(entry.getKey()); } } } return sb.toString(); } You'll notice that this method only processes those panelBoxes that will be in the wrong state and returns those as a comma separated list. This is then processed by the submitDiscloseOverride() method: private void submitDiscloseOverride(String clientIdList, boolean targetDisclosureState) { if (clientIdList != null && clientIdList.length() > 0) { FacesContext fctx = FacesContext.getCurrentInstance(); StringBuilder script = new StringBuilder(); script.append("overrideDiscloseHandler('"); script.append(clientIdList); script.append("',"); script.append(targetDisclosureState); script.append(");"); Service.getRenderKitService(fctx, ExtendedRenderKitService.class).addScript(fctx, script.toString()); } } This method constructs a JavaScript command to call a routine called overrideDiscloseHandler() in a script attached to the page (using the standard <af:resource> tag). That method parses out the list of clientIDs and sends the correct message to each one: function overrideDiscloseHandler(clientIdList, newState) { AdfLogger.LOGGER.logMessage(AdfLogger.INFO, "Disclosure Hander newState " + newState + " Called with: " + clientIdList); //Parse out the list of clientIds var clientIdArray = clientIdList.split(','); for (var i = 0; i < clientIdArray.length; i++){ var panelBox = flipPanel = AdfPage.PAGE.findComponentByAbsoluteId(clientIdArray[i]); if (panelBox.getComponentType() == "oracle.adf.RichPanelBox"){ panelBox.broadcast(new AdfDisclosureEvent(panelBox, newState)); } }  }  So there you go. You can see how, with a few tweaks the same code could be used for other components with disclosure that might suffer from the same problem, although I'd point out that the behavior I'm working around here us usually desirable. You can download the running example (11.1.2.2) from here. 

    Read the article

  • What's up with LDoms: Part 5 - A few Words about Consoles

    - by Stefan Hinker
    Back again to look at a detail of LDom configuration that is often forgotten - the virtual console server. Remember, LDoms are SPARC systems.  As such, each guest will have it's own OBP running.  And to connect to that OBP, the administrator will need a console connection.  Since it's OBP, and not some x86 BIOS, this console will be very serial in nature ;-)  It's really very much like in the good old days, where we had a terminal concentrator where all those serial cables ended up in.  Just like with other components in LDoms, the virtualized solution looks very similar. Every LDom guest requires exactly one console connection.  Envision this similar to the RS-232 port on older SPARC systems.  The LDom framework provides one or more console services that provide access to these connections.  This would be the virtual equivalent of a network terminal server (NTS), where all those serial cables are plugged in.  In the physical world, we'd have a list somewhere, that would tell us which TCP-Port of the NTS was connected to which server.  "ldm list" does just that: root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-cv- UART 16 7680M 0.4% 27d 8h 22m jupiter bound ------ 5002 20 8G mars active -n---- 5000 2 8G 0.5% 55d 14h 10m venus active -n---- 5001 2 8G 0.5% 56d 40m pluto inactive ------ 4 4G The column marked "CONS" tells us, where to reach the console of each domain. In the case of the primary domain, this is actually a (more) physical connection - it's the console connection of the physical system, which is either reachable via the ILOM of that system, or directly via the serial console port on the chassis. All the other guests are reachable through the console service which we created during the inital setup of the system.  Note that pluto does not have a port assigned.  This is because pluto is not yet bound.  (Binding can be viewed very much as the assembly of computer parts - CPU, Memory, disks, network adapters and a serial console cable are all put together when binding the domain.)  Unless we set the port number explicitly, LDoms Manager will do this on a first come, first serve basis.  For just a few domains, this is fine.  For larger deployments, it might be a good idea to assign these port numbers manually using the "ldm set-vcons" command.  However, there is even better magic associated with virtual consoles. You can group several domains into one console group, reachable through one TCP port of the console service.  This can be useful when several groups of administrators are to be given access to different domains, or for other grouping reasons.  Here's an example: root@sun # ldm set-vcons group=planets service=console jupiter root@sun # ldm set-vcons group=planets service=console pluto root@sun # ldm bind jupiter root@sun # ldm bind pluto root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-cv- UART 16 7680M 6.1% 27d 8h 24m jupiter bound ------ 5002 200 8G mars active -n---- 5000 2 8G 0.6% 55d 14h 12m pluto bound ------ 5002 4 4G venus active -n---- 5001 2 8G 0.5% 56d 42m root@sun # telnet localhost 5002 Trying 127.0.0.1... Connected to localhost. Escape character is '^]'. sun-vnts-planets: h, l, c{id}, n{name}, q:l DOMAIN ID DOMAIN NAME DOMAIN STATE 2 jupiter online 3 pluto online sun-vnts-planets: h, l, c{id}, n{name}, q:npluto Connecting to console "pluto" in group "planets" .... Press ~? for control options .. What I did here was add the two domains pluto and jupiter to a new console group called "planets" on the service "console" running in the primary domain.  Simply using a group name will create such a group, if it doesn't already exist.  By default, each domain has its own group, using the domain name as the group name.  The group will be available on port 5002, chosen by LDoms Manager because I didn't specify it.  If I connect to that console group, I will now first be prompted to choose the domain I want to connect to from a little menu. Finally, here's an example how to assign port numbers explicitly: root@sun # ldm set-vcons port=5044 group=pluto service=console pluto root@sun # ldm bind pluto root@sun # ldm list NAME STATE FLAGS CONS VCPU MEMORY UTIL UPTIME primary active -n-cv- UART 16 7680M 3.8% 27d 8h 54m jupiter active -t---- 5002 200 8G 0.5% 30m mars active -n---- 5000 2 8G 0.6% 55d 14h 43m pluto bound ------ 5044 4 4G venus active -n---- 5001 2 8G 0.4% 56d 1h 13m With this, pluto would always be reachable on port 5044 in its own exclusive console group, no matter in which order other domains are bound. Now, you might be wondering why we always have to mention the console service name, "console" in all the examples here.  The simple answer is because there could be more than one such console service.  For all "normal" use, a single console service is absolutely sufficient.  But the system is flexible enough to allow more than that single one, should you need them.  In fact, you could even configure such a console service on a domain other than the primary (or control domain), which would make that domain a real console server.  I actually have a customer who does just that - they want to separate console access from the control domain functionality.  But this is definately a rather sophisticated setup. Something I don't want to go into in this post is access control.  vntsd, which is the daemon providing all these console services, is fully RBAC-aware, and you can configure authorizations for individual users to connect to console groups or individual domain's consoles.  If you can't wait until I get around to security, check out the man page of vntsd. Further reading: The Admin Guide is rather reserved on this subject.  I do recommend to check out the Reference Manual. The manpage for vntsd will discuss all the control sequences as well as the grouping and authorizations mentioned here.

    Read the article

  • Combining Shared Secret and Certificates

    - by Michael Stephenson
    As discussed in the introduction article this walkthrough will explain how you can implement WCF security with the Windows Azure Service Bus to ensure that you can protect your endpoint in the cloud with a shared secret but also combine this with certificates so that you can identify the sender of the message.   Prerequisites As in the previous article before going into the walk through I want to explain a few assumptions about the scenario we are implementing but to keep the article shorter I am not going to walk through all of the steps in how to setup some of this. In the solution we have a simple console application which will represent the client application. There is also the services WCF application which contains the WCF service we will expose via the Windows Azure Service Bus. The WCF Service application in this example was hosted in IIS 7 on Windows 2008 R2 with AppFabric Server installed and configured to auto-start the WCF listening services. I am not going to go through significant detail around the IIS setup because it should not matter in relation to this article however if you want to understand more about how to configure WCF and IIS for such a scenario please refer to the following paper which goes into a lot of detail about how to configure this. The link is: http://tinyurl.com/8s5nwrz   Setting up the Certificates To keep the post and sample simple I am going to use the local computer store for all certificates but this bit is really just the same as setting up certificates for an example where you are using WCF without using Windows Azure Service Bus. In the sample I have included two batch files which you can use to create the sample certificates or remove them. Basically you will end up with: A certificate called PocServerCert in the personal store for the local computer which will be used by the WCF Service component A certificate called PocClientCert in the personal store for the local computer which will be used by the client application A root certificate in the Root store called PocRootCA with its associated revocation list which is the root from which the client and server certificates were created   For the sample Im just using development certificates like you would normally, and you can see exactly how these are configured and placed in the stores from the batch files in the solution using makecert and certmgr.   The Service Component To begin with let's look at the service component and how it can be configured to listen to the service bus using a shared secret but to also accept a username token from the client. In the sample the service component is called Acme.Azure.ServiceBus.Poc.Cert.Services. It has a single service which is the Visual Studio template for a WCF service when you add a new WCF Service Application so we have a service called Service1 with its Echo method. Nothing special so far!.... The next step is to look at the web.config file to see how we have configured the WCF service. In the services section of the WCF configuration you can see I have created my service and I have created a local endpoint which I simply used to do a little bit of diagnostics and to check it was working, but more importantly there is the Windows Azure endpoint which is using the ws2007HttpRelayBinding (note that this should also work just the same if your using netTcpRelayBinding). The key points to note on the above picture are the service behavior called MyServiceBehaviour and the service bus endpoints behavior called MyEndpointBehaviour. We will go into these in more detail later.   The Relay Binding The relay binding for the service has been configured to use the TransportWithMessageCredential security mode. This is the important bit where the transport security really relates to the interaction between the service and listening to the Azure Service Bus and the message credential is where we will use our certificate like we have specified in the message/clientCrentialType attribute. Note also that we have left the relayClientAuthenticationType set to RelayAccessToken. This means that authentication will be made against ACS for accessing the service bus and messages will not be accepted from any sender who has not been authenticated by ACS.   The Endpoint Behaviour In the below picture you can see the endpoint behavior which is configured to use the shared secret client credential for accessing the service bus and also for diagnostic purposes I have included the service registry element.     Hopefully if you are familiar with using Windows Azure Service Bus relay feature the above is very familiar to you and this is a very common setup for this section. There is nothing specific to the username token implementation here. The Service Behaviour Now we come to the bit with most of the certificate stuff in it. When you configure the service behavior I have included the serviceCredentials element and then setup to use the clientCertificate check and also specifying the serviceCertificate with information on how to find the servers certificate in the store.     I have also added a serviceAuthorization section where I will implement my own authorization component to perform additional security checks after the service has validated that the message was signed with a good certificate. I also have the same serviceSecurityAudit configuration to log access to my service. My Authorization Manager The below picture shows you implementation of my authorization manager. WCF will eventually hand off the message to my authorization component before it calls the service code. This is where I can perform some logic to check if the identity is allowed to access resources. In this case I am simple rejecting messages from anyone except the PocClientCertificate.     The Client Now let's take a look at the client side of this solution and how we can configure the client to authenticate against ACS but also send a certificate over to the service component so it can implement additional security checks on-premise. I have a console application and in the program class I want to use the proxy generated with Add Service Reference to send a message via the Azure Service Bus. You can see in my WCF client configuration below I have setup my details for the azure service bus url and am using the ws2007HttpRelayBinding.   Next is my configuration for the relay binding. You can see below I have configured security to use TransportWithMessageCredential so we will flow the token from a certificate with the message and also the RelayAccessToken relayClientAuthenticationType which means the component will validate against ACS before being allowed to access the relay endpoint to send a message.     After the binding we need to configure the endpoint behavior like in the below picture. This contains the normal transportClientEndpointBehaviour to setup the ACS shared secret configuration but we have also configured the clientCertificate to look for the PocClientCert.     Finally below we have the code of the client in the console application which will call the service bus. You can see that we have created our proxy and then made a normal call to a WCF in exactly the normal way but the configuration will jump in and ensure that a token is passed representing the client certificate.     Conclusion As you can see from the above walkthrough it is not too difficult to configure a service to use both a shared secret and certificate based token at the same time. This gives you the power and protection offered by the access control service in the cloud but also the ability to flow additional tokens to the on-premise component for additional security features to be implemented. Sample The sample used in this post is available at the following location: https://s3.amazonaws.com/CSCBlogSamples/Acme.Azure.ServiceBus.Poc.Cert.zip

    Read the article

  • Notes on implementing Visual Studio 2010 Navigate To

    - by cyberycon
    One of the many neat functions added to Visual Studio in VS 2010 was the Navigate To feature. You can find it by clicking Edit, Navigate To, or by using the keyboard shortcut Ctrl, (yes, that's control plus the comma key). This pops up the Navigate To dialog that looks like this: As you type, Navigate To starts searching through a number of different search providers for your term. The entries in the list change as you type, with most providers doing some kind of fuzzy or at least substring matching. If you have C#, C++ or Visual Basic projects in your solution, all symbols defined in those projects are searched. There's also a file search provider, which displays all matching filenames from projects in the current solution as well. And, if you have a Visual Studio package of your own, you can implement a provider too. Micro Focus (where I work) provide the Visual COBOL language inside Visual Studio (http://visualstudiogallery.msdn.microsoft.com/ef9bc810-c133-4581-9429-b01420a9ea40 ), and we wanted to provide this functionality too. This post provides some notes on the things I discovered mainly through trial and error, but also with some kind help from devs inside Microsoft. The expectation of Navigate To is that it searches across the whole solution, not just the current project. So in our case, we wanted to search for all COBOL symbols inside all of our Visual COBOL projects inside the solution. So first of all, here's the Microsoft documentation on Navigate To: http://msdn.microsoft.com/en-us/library/ee844862.aspx . It's the reference information on the Microsoft.VisualStudio.Language.NavigateTo.Interfaces Namespace, and it lists all the interfaces you will need to implement to create your own Navigate To provider. Navigate To uses Visual Studio's latest mechanism for integrating external functionality and services, Managed Extensibility Framework (MEF). MEF components don't require any registration with COM or any other registry entries to be found by Visual Studio. Visual Studio looks in several well-known locations for manifest files (extension.vsixmanifest). It then uses reflection to scan for MEF attributes on classes in the assembly to determine which functionality the assembly provides. MEF itself is actually part of the .NET framework, and you can learn more about it here: http://mef.codeplex.com/. To get started with Visual Studio and MEF you could do worse than look at some of the editor examples on the VSX page http://archive.msdn.microsoft.com/vsx . I've also written a small application to help with switching between development and production MEF assemblies, which you can find on Codeproject: http://www.codeproject.com/KB/miscctrl/MEF_Switch.aspx. The Navigate To interfaces Back to Navigate To, and summarizing the MSDN reference documentation, you need to implement the following interfaces: INavigateToItemProviderFactoryThis is Visual Studio's entry point to your Navigate To implementation, and you must decorate your implementation with the following MEF export attribute: [Export(typeof(INavigateToItemProviderFactory))]  INavigateToItemProvider Your INavigateToItemProviderFactory needs to return your implementation of INavigateToItemProvider. This class implements StartSearch() and StopSearch(). StartSearch() is the guts of your provider, and we'll come back to it in a minute. This object also needs to implement IDisposeable(). INavigateToItemDisplayFactory Your INavigateToItemProvider hands back NavigateToItems to the NavigateTo framework. But to give you good control over what appears in the NavigateTo dialog box, these items will be handed back to your INavigateToItemDisplayFactory, which must create objects implementing INavigateToItemDisplay  INavigateToItemDisplay Each of these objects represents one result in the Navigate To dialog box. As well as providing the description and name of the item, this object also has a NavigateTo() method that should be capable of displaying the item in an editor when invoked. Carrying out the search The lifecycle of your INavigateToItemProvider is the same as that of the Navigate To dialog. This dialog is modal, which makes your implementation a little easier because you know that the user can't be changing things in editors and the IDE while this dialog is up. But the Navigate To dialog DOES NOT run on the main UI thread of the IDE – so you need to be aware of that if you want to interact with editors or other parts of the IDE UI. When the user invokes the Navigate To dialog, your INavigateToItemProvider gets sent a TryCreateNavigateToItemProvider() message. Instantiate your INavigateToItemProvider and hand this back. The sequence diagram below shows what happens next. Your INavigateToItemProvider will get called with StartSearch(), and passed an INavigateToCallback. StartSearch() is an asynchronous request – you must return from this method as soon as possible, and conduct your search on a separate thread. For each match to the search term, instantiate a NavigateToItem object and send it to INavigateToCallback.AddItem(). But as the user types in the Search Terms field, NavigateTo will invoke your StartSearch() method repeatedly with the changing search term. When you receive the next StartSearch() message, you have to abandon your current search, and start a new one. You can't rely on receiving a StopSearch() message every time. Finally, when the Navigate To dialog box is closed by the user, you will get a Dispose() message – that's your cue to abandon any uncompleted searches, and dispose any resources you might be using as part of your search. While you conduct your search invoke INavigateToCallback.ReportProgress() occasionally to provide feedback about how close you are to completing the search. There does not appear to be any particular requirement to how often you invoke ReportProgress(), and you report your progress as the ratio of two integers. In my implementation I report progress in terms of the number of symbols I've searched over the total number of symbols in my dictionary, and send a progress report every 16 symbols. Displaying the Results The Navigate to framework invokes INavigateToItemDisplayProvider.CreateItemDisplay() once for each result you passed to the INavigateToCallback. CreateItemDisplay() is passed the NavigateToItem you handed to the callback, and must return an INavigateToItemDisplay object. NavigateToItem is a sealed class which has a few properties, including the name of the symbol. It also has a Tag property, of type object. This enables you to stash away all the information you will need to create your INavigateToItemDisplay, which must implement an INavigateTo() method to display a symbol in an editor IDE when the user double-clicks an entry in the Navigate To dialog box. Since the tag is of type object, it is up to you, the implementor, to decide what kind of object you store in here, and how it enables the retrieval of other information which is not included in the NavigateToItem properties. Some of the INavigateToItemDisplay properties are self-explanatory, but a couple of them are less obvious: Additional informationThe string you return here is displayed inside brackets on the same line as the Name property. In English locales, Visual Studio includes the preposition "of". If you look at the first line in the Navigate To screenshot at the top of this article, Book_WebRole.Default is the additional information for textBookAuthor, and is the namespace qualified type name the symbol appears in. For procedural COBOL code we display the Program Id as the additional information DescriptionItemsYou can use this property to return any textual description you want about the item currently selected. You return a collection of DescriptionItem objects, each of which has a category and description collection of DescriptionRun objects. A DescriptionRun enables you to specify some text, and optional formatting, so you have some control over the appearance of the displayed text. The DescriptionItems property is displayed at the bottom of the Navigate To dialog box, with the Categories on the left and the Descriptions on the right. The Visual COBOL implementation uses it to display more information about the location of an item, making it easier for the user to know disambiguate duplicate names (something there can be a lot of in large COBOL applications). Summary I hope this article is useful for anyone implementing Navigate To. It is a fantastic navigation feature that Microsoft have added to Visual Studio, but at the moment there still don't seem to be any examples on how to implement it, and the reference information on MSDN is a little brief for anyone attempting an implementation.

    Read the article

< Previous Page | 202 203 204 205 206 207 208 209 210 211 212 213  | Next Page >