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  • Ubuntu: Failure to login with multiple video adapters

    - by tsilb
    Forgive my ignorance, for I am a complete linux noob. I have a computer with three video cards and six monitors. Works great on Windows. Trying to get it to run Ubuntu as well. It loads fine when I have it configured to run on one adapter; detects both screens, runs ok. But I want to turn the other 4 monitors on and run the whole thing as one extended desktop (one session, etc). So I downloaded and installed the newest ATI driver for Linux, which seems to work, kinda. I ran this to set up the screens: aticonfig --adapter=all --initial -f Now when I boot, Ubuntu seems to turn on all the screens (3 viewports, each with two cloned displays from what I can tell). When I enter my login info OR move the mouse off the main screen, the screens freeze and the kbd/ms become unresponsive. aticonfig generated xorg.conf included below. Have tried the following: aticonfig -initial -f - works, but only detects the primary adapter and 2 screens aticccle - Tells me I have to reboot after enabling the other cards. Then goes into above described freezing state. aticonfig --adapter=all --initial -f - see above Manually editing xorg.conf file with my limited knowledge - Was able to get two adapters running, but only the second adapter initialized while the primary stopped at the Ubuntu boot screen. Was unable to see the login prompt. Froze after I logged in blindly (was able to hear the login sound). Using generic "radeon" driver instead of ATI Proprietary driver with the above init attempts Toggling xinerama Various combinations of the above Hardware: Intel Core 2 Quad q6600 8GB DDR2 (3x) ATI Radeon HD 4680 5 monitors (21W, 21W, 22W Portrait, 22W Portrait, 19")and an HDTV (26"W, HDMI) in a horizontal arrangement I know next to nothing about Linux/Ubuntu aside from basic filesystem navigation, editing text files, and accessing my local and networked Windows stores and shares. Basically this is the most advanced thing I've had to do. I installed today. Please advise how to make this configuration work. my xorg.conf: Section "ServerLayout" Identifier "Layout0" Screen 0 "aticonfig-Screen[0]-0" 0 0 Screen "aticonfig-Screen[1]-0" RightOf "aticonfig-Screen[0]-0" Screen "aticonfig-Screen[2]-0" RightOf "aticonfig-Screen[1]-0" Option "RenderAccel" "true" Option "AllowGLXWithComposite" "true" EndSection Section "Files" EndSection Section "Module" EndSection Section "ServerFlags" Option "Xinerama" "0" EndSection Section "Monitor" Identifier "aticonfig-Monitor[0]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Monitor" Identifier "aticonfig-Monitor[1]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Monitor" Identifier "aticonfig-Monitor[2]-0" Option "VendorName" "ATI Proprietary Driver" Option "ModelName" "Generic Autodetecting Monitor" Option "DPMS" "true" EndSection Section "Device" Identifier "aticonfig-Device[0]-0" Driver "fglrx" BusID "PCI:1:0:0" EndSection Section "Device" Identifier "aticonfig-Device[1]-0" Driver "fglrx" BusID "PCI:3:0:0" EndSection Section "Device" Identifier "aticonfig-Device[2]-0" Driver "fglrx" BusID "PCI:4:0:0" EndSection Section "Screen" Identifier "aticonfig-Screen[0]-0" Device "aticonfig-Device[0]-0" Monitor "aticonfig-Monitor[0]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection Section "Screen" Identifier "aticonfig-Screen[1]-0" Device "aticonfig-Device[1]-0" Monitor "aticonfig-Monitor[1]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection Section "Screen" Identifier "aticonfig-Screen[2]-0" Device "aticonfig-Device[2]-0" Monitor "aticonfig-Monitor[2]-0" DefaultDepth 24 SubSection "Display" Viewport 0 0 Depth 24 EndSubSection EndSection

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  • SUSE EC2 Problem - zypper - Permission denied

    - by phuu
    Hi. I'm trying to use zypper to install gcc on my Amazon EC2 instance running SUSE.When I try:zypper in gcc I get: Retrieving repository 'SLE11-SDK-SP1' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLE11-SDK-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1' metadata [error] Repository 'SLE11-SDK-SP1' is invalid. Can't provide /media.1/media : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1' because of the above error. Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [|] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLE11-SDK-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): i Retrieving repository 'SLE11-SDK-SP1-Updates' metadata [error] Repository 'SLE11-SDK-SP1-Updates' is invalid. Can't provide /repodata/repomd.xml : User-requested skipping of a file Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLE11-SDK-SP1-Updates' because of the above error. Retrieving repository 'SLES11-Extras' metadata [/] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): r Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-Extras/sle-11-i586/repodata/repomd.xml' denied. Abort, retry, ignore? [a/r/i/?] (a): zypper in gcc Invalid answer 'zypper in gcc'. [a/r/i/?] (a): a Retrieving repository 'SLES11-Extras' metadata [error] Repository 'SLES11-Extras' is invalid. Can't provide /repodata/repomd.xml : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-Extras' because of the above error. Retrieving repository 'SLES11-SP1' metadata [-] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/install/SLES11-SP1/sle-11-i586/media.1/media' denied. Abort, retry, ignore? [a/r/i/?] (a): a Retrieving repository 'SLES11-SP1' metadata [error] Repository 'SLES11-SP1' is invalid. Can't provide /media.1/media : Please check if the URIs defined for this repository are pointing to a valid repository. Warning: Disabling repository 'SLES11-SP1' because of the above error. Retrieving repository 'SLES11-SP1-Updates' metadata [] Permission to access 'http://eu-west-1-ec2-update.susecloud.net/repo/update/SLES11-SP1-Updates/sle-11-i586/repodata/repomd.xml' denied. I've search for the problem and this thread came up, but offered no solutions.I've triedsces-activate. Am I doing something wrong? I should say I'm very new to this, and I admit I don't really know what I'm doing, but I'm trying to learn about setting up and running a server and so I thought I'd throw myself in at the deep(ish) end. Thanks for reading.

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  • How to setup linux permissions for the WWW folder?

    - by Xeoncross
    Updated Summery The /var/www directory is owned by root:root which means that no one can use it and it's entirely useless. Since we all want a web server that actually works (and no-one should be logging in as "root"), then we need to fix this. Only two entities need access. PHP/Perl/Ruby/Python all need access to the folders and files since they create many of them (i.e. /uploads/). These scripting languages should be running under nginx or apache (or even some other thing like FastCGI for PHP). The developers How do they get access? I know that someone, somewhere has done this before. With however-many billions of websites out there you would think that there would be more information on this topic. I know that 777 is full read/write/execute permission for owner/group/other. So this doesn't seem to be needed as it leaves random users full permissions. What permissions are need to be used on /var/www so that... Source control like git or svn Users in a group like "websites" (or even added to "www-data") Servers like apache or lighthttpd And PHP/Perl/Ruby can all read, create, and run files (and directories) there? If I'm correct, Ruby and PHP scripts are not "executed" directly - but passed to an interpreter. So there is no need for execute permission on files in /var/www...? Therefore, it seems like the correct permission would be chmod -R 1660 which would make all files shareable by these four entities all files non-executable by mistake block everyone else from the directory entirely set the permission mode to "sticky" for all future files Is this correct? Update: I just realized that files and directories might need different permissions - I was talking about files above so i'm not sure what the directory permissions would need to be. Update 2: The folder structure of /var/www changes drastically as one of the four entities above are always adding (and sometimes removing) folders and sub folders many levels deep. They also create and remove files that the other 3 entities might need read/write access to. Therefore, the permissions need to do the four things above for both files and directories. Since non of them should need execute permission (see question about ruby/php above) I would assume that rw-rw-r-- permission would be all that is needed and completely safe since these four entities are run by trusted personal (see #2) and all other users on the system only have read access. Update 3: This is for personal development machines and private company servers. No random "web customers" like a shared host. Update 4: This article by slicehost seems to be the best at explaining what is needed to setup permissions for your www folder. However, I'm not sure what user or group apache/nginx with PHP OR svn/git run as and how to change them. Update 5: I have (I think) finally found a way to get this all to work (answer below). However, I don't know if this is the correct and SECURE way to do this. Therefore I have started a bounty. The person that has the best method of securing and managing the www directory wins.

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  • value types in the vm

    - by john.rose
    value types in the vm p.p1 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times} p.p2 {margin: 0.0px 0.0px 14.0px 0.0px; font: 14.0px Times} p.p3 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times} p.p4 {margin: 0.0px 0.0px 15.0px 0.0px; font: 14.0px Times} p.p5 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Courier} p.p6 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Courier; min-height: 17.0px} p.p7 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times; min-height: 18.0px} p.p8 {margin: 0.0px 0.0px 0.0px 36.0px; text-indent: -36.0px; font: 14.0px Times; min-height: 18.0px} p.p9 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times; min-height: 18.0px} p.p10 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times; color: #000000} li.li1 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times} li.li7 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times; min-height: 18.0px} span.s1 {font: 14.0px Courier} span.s2 {color: #000000} span.s3 {font: 14.0px Courier; color: #000000} ol.ol1 {list-style-type: decimal} Or, enduring values for a changing world. Introduction A value type is a data type which, generally speaking, is designed for being passed by value in and out of methods, and stored by value in data structures. The only value types which the Java language directly supports are the eight primitive types. Java indirectly and approximately supports value types, if they are implemented in terms of classes. For example, both Integer and String may be viewed as value types, especially if their usage is restricted to avoid operations appropriate to Object. In this note, we propose a definition of value types in terms of a design pattern for Java classes, accompanied by a set of usage restrictions. We also sketch the relation of such value types to tuple types (which are a JVM-level notion), and point out JVM optimizations that can apply to value types. This note is a thought experiment to extend the JVM’s performance model in support of value types. The demonstration has two phases.  Initially the extension can simply use design patterns, within the current bytecode architecture, and in today’s Java language. But if the performance model is to be realized in practice, it will probably require new JVM bytecode features, changes to the Java language, or both.  We will look at a few possibilities for these new features. An Axiom of Value In the context of the JVM, a value type is a data type equipped with construction, assignment, and equality operations, and a set of typed components, such that, whenever two variables of the value type produce equal corresponding values for their components, the values of the two variables cannot be distinguished by any JVM operation. Here are some corollaries: A value type is immutable, since otherwise a copy could be constructed and the original could be modified in one of its components, allowing the copies to be distinguished. Changing the component of a value type requires construction of a new value. The equals and hashCode operations are strictly component-wise. If a value type is represented by a JVM reference, that reference cannot be successfully synchronized on, and cannot be usefully compared for reference equality. A value type can be viewed in terms of what it doesn’t do. We can say that a value type omits all value-unsafe operations, which could violate the constraints on value types.  These operations, which are ordinarily allowed for Java object types, are pointer equality comparison (the acmp instruction), synchronization (the monitor instructions), all the wait and notify methods of class Object, and non-trivial finalize methods. The clone method is also value-unsafe, although for value types it could be treated as the identity function. Finally, and most importantly, any side effect on an object (however visible) also counts as an value-unsafe operation. A value type may have methods, but such methods must not change the components of the value. It is reasonable and useful to define methods like toString, equals, and hashCode on value types, and also methods which are specifically valuable to users of the value type. Representations of Value Value types have two natural representations in the JVM, unboxed and boxed. An unboxed value consists of the components, as simple variables. For example, the complex number x=(1+2i), in rectangular coordinate form, may be represented in unboxed form by the following pair of variables: /*Complex x = Complex.valueOf(1.0, 2.0):*/ double x_re = 1.0, x_im = 2.0; These variables might be locals, parameters, or fields. Their association as components of a single value is not defined to the JVM. Here is a sample computation which computes the norm of the difference between two complex numbers: double distance(/*Complex x:*/ double x_re, double x_im,         /*Complex y:*/ double y_re, double y_im) {     /*Complex z = x.minus(y):*/     double z_re = x_re - y_re, z_im = x_im - y_im;     /*return z.abs():*/     return Math.sqrt(z_re*z_re + z_im*z_im); } A boxed representation groups component values under a single object reference. The reference is to a ‘wrapper class’ that carries the component values in its fields. (A primitive type can naturally be equated with a trivial value type with just one component of that type. In that view, the wrapper class Integer can serve as a boxed representation of value type int.) The unboxed representation of complex numbers is practical for many uses, but it fails to cover several major use cases: return values, array elements, and generic APIs. The two components of a complex number cannot be directly returned from a Java function, since Java does not support multiple return values. The same story applies to array elements: Java has no ’array of structs’ feature. (Double-length arrays are a possible workaround for complex numbers, but not for value types with heterogeneous components.) By generic APIs I mean both those which use generic types, like Arrays.asList and those which have special case support for primitive types, like String.valueOf and PrintStream.println. Those APIs do not support unboxed values, and offer some problems to boxed values. Any ’real’ JVM type should have a story for returns, arrays, and API interoperability. The basic problem here is that value types fall between primitive types and object types. Value types are clearly more complex than primitive types, and object types are slightly too complicated. Objects are a little bit dangerous to use as value carriers, since object references can be compared for pointer equality, and can be synchronized on. Also, as many Java programmers have observed, there is often a performance cost to using wrapper objects, even on modern JVMs. Even so, wrapper classes are a good starting point for talking about value types. If there were a set of structural rules and restrictions which would prevent value-unsafe operations on value types, wrapper classes would provide a good notation for defining value types. This note attempts to define such rules and restrictions. Let’s Start Coding Now it is time to look at some real code. Here is a definition, written in Java, of a complex number value type. @ValueSafe public final class Complex implements java.io.Serializable {     // immutable component structure:     public final double re, im;     private Complex(double re, double im) {         this.re = re; this.im = im;     }     // interoperability methods:     public String toString() { return "Complex("+re+","+im+")"; }     public List<Double> asList() { return Arrays.asList(re, im); }     public boolean equals(Complex c) {         return re == c.re && im == c.im;     }     public boolean equals(@ValueSafe Object x) {         return x instanceof Complex && equals((Complex) x);     }     public int hashCode() {         return 31*Double.valueOf(re).hashCode()                 + Double.valueOf(im).hashCode();     }     // factory methods:     public static Complex valueOf(double re, double im) {         return new Complex(re, im);     }     public Complex changeRe(double re2) { return valueOf(re2, im); }     public Complex changeIm(double im2) { return valueOf(re, im2); }     public static Complex cast(@ValueSafe Object x) {         return x == null ? ZERO : (Complex) x;     }     // utility methods and constants:     public Complex plus(Complex c)  { return new Complex(re+c.re, im+c.im); }     public Complex minus(Complex c) { return new Complex(re-c.re, im-c.im); }     public double abs() { return Math.sqrt(re*re + im*im); }     public static final Complex PI = valueOf(Math.PI, 0.0);     public static final Complex ZERO = valueOf(0.0, 0.0); } This is not a minimal definition, because it includes some utility methods and other optional parts.  The essential elements are as follows: The class is marked as a value type with an annotation. The class is final, because it does not make sense to create subclasses of value types. The fields of the class are all non-private and final.  (I.e., the type is immutable and structurally transparent.) From the supertype Object, all public non-final methods are overridden. The constructor is private. Beyond these bare essentials, we can observe the following features in this example, which are likely to be typical of all value types: One or more factory methods are responsible for value creation, including a component-wise valueOf method. There are utility methods for complex arithmetic and instance creation, such as plus and changeIm. There are static utility constants, such as PI. The type is serializable, using the default mechanisms. There are methods for converting to and from dynamically typed references, such as asList and cast. The Rules In order to use value types properly, the programmer must avoid value-unsafe operations.  A helpful Java compiler should issue errors (or at least warnings) for code which provably applies value-unsafe operations, and should issue warnings for code which might be correct but does not provably avoid value-unsafe operations.  No such compilers exist today, but to simplify our account here, we will pretend that they do exist. A value-safe type is any class, interface, or type parameter marked with the @ValueSafe annotation, or any subtype of a value-safe type.  If a value-safe class is marked final, it is in fact a value type.  All other value-safe classes must be abstract.  The non-static fields of a value class must be non-public and final, and all its constructors must be private. Under the above rules, a standard interface could be helpful to define value types like Complex.  Here is an example: @ValueSafe public interface ValueType extends java.io.Serializable {     // All methods listed here must get redefined.     // Definitions must be value-safe, which means     // they may depend on component values only.     List<? extends Object> asList();     int hashCode();     boolean equals(@ValueSafe Object c);     String toString(); } //@ValueSafe inherited from supertype: public final class Complex implements ValueType { … The main advantage of such a conventional interface is that (unlike an annotation) it is reified in the runtime type system.  It could appear as an element type or parameter bound, for facilities which are designed to work on value types only.  More broadly, it might assist the JVM to perform dynamic enforcement of the rules for value types. Besides types, the annotation @ValueSafe can mark fields, parameters, local variables, and methods.  (This is redundant when the type is also value-safe, but may be useful when the type is Object or another supertype of a value type.)  Working forward from these annotations, an expression E is defined as value-safe if it satisfies one or more of the following: The type of E is a value-safe type. E names a field, parameter, or local variable whose declaration is marked @ValueSafe. E is a call to a method whose declaration is marked @ValueSafe. E is an assignment to a value-safe variable, field reference, or array reference. E is a cast to a value-safe type from a value-safe expression. E is a conditional expression E0 ? E1 : E2, and both E1 and E2 are value-safe. Assignments to value-safe expressions and initializations of value-safe names must take their values from value-safe expressions. A value-safe expression may not be the subject of a value-unsafe operation.  In particular, it cannot be synchronized on, nor can it be compared with the “==” operator, not even with a null or with another value-safe type. In a program where all of these rules are followed, no value-type value will be subject to a value-unsafe operation.  Thus, the prime axiom of value types will be satisfied, that no two value type will be distinguishable as long as their component values are equal. More Code To illustrate these rules, here are some usage examples for Complex: Complex pi = Complex.valueOf(Math.PI, 0); Complex zero = pi.changeRe(0);  //zero = pi; zero.re = 0; ValueType vtype = pi; @SuppressWarnings("value-unsafe")   Object obj = pi; @ValueSafe Object obj2 = pi; obj2 = new Object();  // ok List<Complex> clist = new ArrayList<Complex>(); clist.add(pi);  // (ok assuming List.add param is @ValueSafe) List<ValueType> vlist = new ArrayList<ValueType>(); vlist.add(pi);  // (ok) List<Object> olist = new ArrayList<Object>(); olist.add(pi);  // warning: "value-unsafe" boolean z = pi.equals(zero); boolean z1 = (pi == zero);  // error: reference comparison on value type boolean z2 = (pi == null);  // error: reference comparison on value type boolean z3 = (pi == obj2);  // error: reference comparison on value type synchronized (pi) { }  // error: synch of value, unpredictable result synchronized (obj2) { }  // unpredictable result Complex qq = pi; qq = null;  // possible NPE; warning: “null-unsafe" qq = (Complex) obj;  // warning: “null-unsafe" qq = Complex.cast(obj);  // OK @SuppressWarnings("null-unsafe")   Complex empty = null;  // possible NPE qq = empty;  // possible NPE (null pollution) The Payoffs It follows from this that either the JVM or the java compiler can replace boxed value-type values with unboxed ones, without affecting normal computations.  Fields and variables of value types can be split into their unboxed components.  Non-static methods on value types can be transformed into static methods which take the components as value parameters. Some common questions arise around this point in any discussion of value types. Why burden the programmer with all these extra rules?  Why not detect programs automagically and perform unboxing transparently?  The answer is that it is easy to break the rules accidently unless they are agreed to by the programmer and enforced.  Automatic unboxing optimizations are tantalizing but (so far) unreachable ideal.  In the current state of the art, it is possible exhibit benchmarks in which automatic unboxing provides the desired effects, but it is not possible to provide a JVM with a performance model that assures the programmer when unboxing will occur.  This is why I’m writing this note, to enlist help from, and provide assurances to, the programmer.  Basically, I’m shooting for a good set of user-supplied “pragmas” to frame the desired optimization. Again, the important thing is that the unboxing must be done reliably, or else programmers will have no reason to work with the extra complexity of the value-safety rules.  There must be a reasonably stable performance model, wherein using a value type has approximately the same performance characteristics as writing the unboxed components as separate Java variables. There are some rough corners to the present scheme.  Since Java fields and array elements are initialized to null, value-type computations which incorporate uninitialized variables can produce null pointer exceptions.  One workaround for this is to require such variables to be null-tested, and the result replaced with a suitable all-zero value of the value type.  That is what the “cast” method does above. Generically typed APIs like List<T> will continue to manipulate boxed values always, at least until we figure out how to do reification of generic type instances.  Use of such APIs will elicit warnings until their type parameters (and/or relevant members) are annotated or typed as value-safe.  Retrofitting List<T> is likely to expose flaws in the present scheme, which we will need to engineer around.  Here are a couple of first approaches: public interface java.util.List<@ValueSafe T> extends Collection<T> { … public interface java.util.List<T extends Object|ValueType> extends Collection<T> { … (The second approach would require disjunctive types, in which value-safety is “contagious” from the constituent types.) With more transformations, the return value types of methods can also be unboxed.  This may require significant bytecode-level transformations, and would work best in the presence of a bytecode representation for multiple value groups, which I have proposed elsewhere under the title “Tuples in the VM”. But for starters, the JVM can apply this transformation under the covers, to internally compiled methods.  This would give a way to express multiple return values and structured return values, which is a significant pain-point for Java programmers, especially those who work with low-level structure types favored by modern vector and graphics processors.  The lack of multiple return values has a strong distorting effect on many Java APIs. Even if the JVM fails to unbox a value, there is still potential benefit to the value type.  Clustered computing systems something have copy operations (serialization or something similar) which apply implicitly to command operands.  When copying JVM objects, it is extremely helpful to know when an object’s identity is important or not.  If an object reference is a copied operand, the system may have to create a proxy handle which points back to the original object, so that side effects are visible.  Proxies must be managed carefully, and this can be expensive.  On the other hand, value types are exactly those types which a JVM can “copy and forget” with no downside. Array types are crucial to bulk data interfaces.  (As data sizes and rates increase, bulk data becomes more important than scalar data, so arrays are definitely accompanying us into the future of computing.)  Value types are very helpful for adding structure to bulk data, so a successful value type mechanism will make it easier for us to express richer forms of bulk data. Unboxing arrays (i.e., arrays containing unboxed values) will provide better cache and memory density, and more direct data movement within clustered or heterogeneous computing systems.  They require the deepest transformations, relative to today’s JVM.  There is an impedance mismatch between value-type arrays and Java’s covariant array typing, so compromises will need to be struck with existing Java semantics.  It is probably worth the effort, since arrays of unboxed value types are inherently more memory-efficient than standard Java arrays, which rely on dependent pointer chains. It may be sufficient to extend the “value-safe” concept to array declarations, and allow low-level transformations to change value-safe array declarations from the standard boxed form into an unboxed tuple-based form.  Such value-safe arrays would not be convertible to Object[] arrays.  Certain connection points, such as Arrays.copyOf and System.arraycopy might need additional input/output combinations, to allow smooth conversion between arrays with boxed and unboxed elements. Alternatively, the correct solution may have to wait until we have enough reification of generic types, and enough operator overloading, to enable an overhaul of Java arrays. Implicit Method Definitions The example of class Complex above may be unattractively complex.  I believe most or all of the elements of the example class are required by the logic of value types. If this is true, a programmer who writes a value type will have to write lots of error-prone boilerplate code.  On the other hand, I think nearly all of the code (except for the domain-specific parts like plus and minus) can be implicitly generated. Java has a rule for implicitly defining a class’s constructor, if no it defines no constructors explicitly.  Likewise, there are rules for providing default access modifiers for interface members.  Because of the highly regular structure of value types, it might be reasonable to perform similar implicit transformations on value types.  Here’s an example of a “highly implicit” definition of a complex number type: public class Complex implements ValueType {  // implicitly final     public double re, im;  // implicitly public final     //implicit methods are defined elementwise from te fields:     //  toString, asList, equals(2), hashCode, valueOf, cast     //optionally, explicit methods (plus, abs, etc.) would go here } In other words, with the right defaults, a simple value type definition can be a one-liner.  The observant reader will have noticed the similarities (and suitable differences) between the explicit methods above and the corresponding methods for List<T>. Another way to abbreviate such a class would be to make an annotation the primary trigger of the functionality, and to add the interface(s) implicitly: public @ValueType class Complex { … // implicitly final, implements ValueType (But to me it seems better to communicate the “magic” via an interface, even if it is rooted in an annotation.) Implicitly Defined Value Types So far we have been working with nominal value types, which is to say that the sequence of typed components is associated with a name and additional methods that convey the intention of the programmer.  A simple ordered pair of floating point numbers can be variously interpreted as (to name a few possibilities) a rectangular or polar complex number or Cartesian point.  The name and the methods convey the intended meaning. But what if we need a truly simple ordered pair of floating point numbers, without any further conceptual baggage?  Perhaps we are writing a method (like “divideAndRemainder”) which naturally returns a pair of numbers instead of a single number.  Wrapping the pair of numbers in a nominal type (like “QuotientAndRemainder”) makes as little sense as wrapping a single return value in a nominal type (like “Quotient”).  What we need here are structural value types commonly known as tuples. For the present discussion, let us assign a conventional, JVM-friendly name to tuples, roughly as follows: public class java.lang.tuple.$DD extends java.lang.tuple.Tuple {      double $1, $2; } Here the component names are fixed and all the required methods are defined implicitly.  The supertype is an abstract class which has suitable shared declarations.  The name itself mentions a JVM-style method parameter descriptor, which may be “cracked” to determine the number and types of the component fields. The odd thing about such a tuple type (and structural types in general) is it must be instantiated lazily, in response to linkage requests from one or more classes that need it.  The JVM and/or its class loaders must be prepared to spin a tuple type on demand, given a simple name reference, $xyz, where the xyz is cracked into a series of component types.  (Specifics of naming and name mangling need some tasteful engineering.) Tuples also seem to demand, even more than nominal types, some support from the language.  (This is probably because notations for non-nominal types work best as combinations of punctuation and type names, rather than named constructors like Function3 or Tuple2.)  At a minimum, languages with tuples usually (I think) have some sort of simple bracket notation for creating tuples, and a corresponding pattern-matching syntax (or “destructuring bind”) for taking tuples apart, at least when they are parameter lists.  Designing such a syntax is no simple thing, because it ought to play well with nominal value types, and also with pre-existing Java features, such as method parameter lists, implicit conversions, generic types, and reflection.  That is a task for another day. Other Use Cases Besides complex numbers and simple tuples there are many use cases for value types.  Many tuple-like types have natural value-type representations. These include rational numbers, point locations and pixel colors, and various kinds of dates and addresses. Other types have a variable-length ‘tail’ of internal values. The most common example of this is String, which is (mathematically) a sequence of UTF-16 character values. Similarly, bit vectors, multiple-precision numbers, and polynomials are composed of sequences of values. Such types include, in their representation, a reference to a variable-sized data structure (often an array) which (somehow) represents the sequence of values. The value type may also include ’header’ information. Variable-sized values often have a length distribution which favors short lengths. In that case, the design of the value type can make the first few values in the sequence be direct ’header’ fields of the value type. In the common case where the header is enough to represent the whole value, the tail can be a shared null value, or even just a null reference. Note that the tail need not be an immutable object, as long as the header type encapsulates it well enough. This is the case with String, where the tail is a mutable (but never mutated) character array. Field types and their order must be a globally visible part of the API.  The structure of the value type must be transparent enough to have a globally consistent unboxed representation, so that all callers and callees agree about the type and order of components  that appear as parameters, return types, and array elements.  This is a trade-off between efficiency and encapsulation, which is forced on us when we remove an indirection enjoyed by boxed representations.  A JVM-only transformation would not care about such visibility, but a bytecode transformation would need to take care that (say) the components of complex numbers would not get swapped after a redefinition of Complex and a partial recompile.  Perhaps constant pool references to value types need to declare the field order as assumed by each API user. This brings up the delicate status of private fields in a value type.  It must always be possible to load, store, and copy value types as coordinated groups, and the JVM performs those movements by moving individual scalar values between locals and stack.  If a component field is not public, what is to prevent hostile code from plucking it out of the tuple using a rogue aload or astore instruction?  Nothing but the verifier, so we may need to give it more smarts, so that it treats value types as inseparable groups of stack slots or locals (something like long or double). My initial thought was to make the fields always public, which would make the security problem moot.  But public is not always the right answer; consider the case of String, where the underlying mutable character array must be encapsulated to prevent security holes.  I believe we can win back both sides of the tradeoff, by training the verifier never to split up the components in an unboxed value.  Just as the verifier encapsulates the two halves of a 64-bit primitive, it can encapsulate the the header and body of an unboxed String, so that no code other than that of class String itself can take apart the values. Similar to String, we could build an efficient multi-precision decimal type along these lines: public final class DecimalValue extends ValueType {     protected final long header;     protected private final BigInteger digits;     public DecimalValue valueOf(int value, int scale) {         assert(scale >= 0);         return new DecimalValue(((long)value << 32) + scale, null);     }     public DecimalValue valueOf(long value, int scale) {         if (value == (int) value)             return valueOf((int)value, scale);         return new DecimalValue(-scale, new BigInteger(value));     } } Values of this type would be passed between methods as two machine words. Small values (those with a significand which fits into 32 bits) would be represented without any heap data at all, unless the DecimalValue itself were boxed. (Note the tension between encapsulation and unboxing in this case.  It would be better if the header and digits fields were private, but depending on where the unboxing information must “leak”, it is probably safer to make a public revelation of the internal structure.) Note that, although an array of Complex can be faked with a double-length array of double, there is no easy way to fake an array of unboxed DecimalValues.  (Either an array of boxed values or a transposed pair of homogeneous arrays would be reasonable fallbacks, in a current JVM.)  Getting the full benefit of unboxing and arrays will require some new JVM magic. Although the JVM emphasizes portability, system dependent code will benefit from using machine-level types larger than 64 bits.  For example, the back end of a linear algebra package might benefit from value types like Float4 which map to stock vector types.  This is probably only worthwhile if the unboxing arrays can be packed with such values. More Daydreams A more finely-divided design for dynamic enforcement of value safety could feature separate marker interfaces for each invariant.  An empty marker interface Unsynchronizable could cause suitable exceptions for monitor instructions on objects in marked classes.  More radically, a Interchangeable marker interface could cause JVM primitives that are sensitive to object identity to raise exceptions; the strangest result would be that the acmp instruction would have to be specified as raising an exception. @ValueSafe public interface ValueType extends java.io.Serializable,         Unsynchronizable, Interchangeable { … public class Complex implements ValueType {     // inherits Serializable, Unsynchronizable, Interchangeable, @ValueSafe     … It seems possible that Integer and the other wrapper types could be retro-fitted as value-safe types.  This is a major change, since wrapper objects would be unsynchronizable and their references interchangeable.  It is likely that code which violates value-safety for wrapper types exists but is uncommon.  It is less plausible to retro-fit String, since the prominent operation String.intern is often used with value-unsafe code. We should also reconsider the distinction between boxed and unboxed values in code.  The design presented above obscures that distinction.  As another thought experiment, we could imagine making a first class distinction in the type system between boxed and unboxed representations.  Since only primitive types are named with a lower-case initial letter, we could define that the capitalized version of a value type name always refers to the boxed representation, while the initial lower-case variant always refers to boxed.  For example: complex pi = complex.valueOf(Math.PI, 0); Complex boxPi = pi;  // convert to boxed myList.add(boxPi); complex z = myList.get(0);  // unbox Such a convention could perhaps absorb the current difference between int and Integer, double and Double. It might also allow the programmer to express a helpful distinction among array types. As said above, array types are crucial to bulk data interfaces, but are limited in the JVM.  Extending arrays beyond the present limitations is worth thinking about; for example, the Maxine JVM implementation has a hybrid object/array type.  Something like this which can also accommodate value type components seems worthwhile.  On the other hand, does it make sense for value types to contain short arrays?  And why should random-access arrays be the end of our design process, when bulk data is often sequentially accessed, and it might make sense to have heterogeneous streams of data as the natural “jumbo” data structure.  These considerations must wait for another day and another note. More Work It seems to me that a good sequence for introducing such value types would be as follows: Add the value-safety restrictions to an experimental version of javac. Code some sample applications with value types, including Complex and DecimalValue. Create an experimental JVM which internally unboxes value types but does not require new bytecodes to do so.  Ensure the feasibility of the performance model for the sample applications. Add tuple-like bytecodes (with or without generic type reification) to a major revision of the JVM, and teach the Java compiler to switch in the new bytecodes without code changes. A staggered roll-out like this would decouple language changes from bytecode changes, which is always a convenient thing. A similar investigation should be applied (concurrently) to array types.  In this case, it seems to me that the starting point is in the JVM: Add an experimental unboxing array data structure to a production JVM, perhaps along the lines of Maxine hybrids.  No bytecode or language support is required at first; everything can be done with encapsulated unsafe operations and/or method handles. Create an experimental JVM which internally unboxes value types but does not require new bytecodes to do so.  Ensure the feasibility of the performance model for the sample applications. Add tuple-like bytecodes (with or without generic type reification) to a major revision of the JVM, and teach the Java compiler to switch in the new bytecodes without code changes. That’s enough musing me for now.  Back to work!

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  • Flow-Design Cheat Sheet &ndash; Part I, Notation

    - by Ralf Westphal
    You want to avoid the pitfalls of object oriented design? Then this is the right place to start. Use Flow-Oriented Analysis (FOA) and –Design (FOD or just FD for Flow-Design) to understand a problem domain and design a software solution. Flow-Orientation as described here is related to Flow-Based Programming, Event-Based Programming, Business Process Modelling, and even Event-Driven Architectures. But even though “thinking in flows” is not new, I found it helpful to deviate from those precursors for several reasons. Some aim at too big systems for the average programmer, some are concerned with only asynchronous processing, some are even not very much concerned with programming at all. What I was looking for was a design method to help in software projects of any size, be they large or tiny, involing synchronous or asynchronous processing, being local or distributed, running on the web or on the desktop or on a smartphone. That´s why I took ideas from all of the above sources and some additional and came up with Event-Based Components which later got repositioned and renamed to Flow-Design. In the meantime this has generated some discussion (in the German developer community) and several teams have started to work with Flow-Design. Also I´ve conducted quite some trainings using Flow-Orientation for design. The results are very promising. Developers find it much easier to design software using Flow-Orientation than OOAD-based object orientation. Since Flow-Orientation is moving fast and is not covered completely by a single source like a book, demand has increased for at least an overview of the current state of its notation. This page is trying to answer this demand by briefly introducing/describing every notational element as well as their translation into C# source code. Take this as a cheat sheet to put next to your whiteboard when designing software. However, please do not expect any explanation as to the reasons behind Flow-Design elements. Details on why Flow-Design at all and why in this specific way you´ll find in the literature covering the topic. Here´s a resource page on Flow-Design/Event-Based Components, if you´re able to read German. Notation Connected Functional Units The basic element of any FOD are functional units (FU): Think of FUs as some kind of software code block processing data. For the moment forget about classes, methods, “components”, assemblies or whatever. See a FU as an abstract piece of code. Software then consists of just collaborating FUs. I´m using circles/ellipses to draw FUs. But if you like, use rectangles. Whatever suites your whiteboard needs best.   The purpose of FUs is to process input and produce output. FUs are transformational. However, FUs are not called and do not call other FUs. There is no dependency between FUs. Data just flows into a FU (input) and out of it (output). From where and where to is of no concern to a FU.   This way FUs can be concatenated in arbitrary ways:   Each FU can accept input from many sources and produce output for many sinks:   Flows Connected FUs form a flow with a start and an end. Data is entering a flow at a source, and it´s leaving it through a sink. Think of sources and sinks as special FUs which conntect wires to the environment of a network of FUs.   Wiring Details Data is flowing into/out of FUs through wires. This is to allude to electrical engineering which since long has been working with composable parts. Wires are attached to FUs usings pins. They are the entry/exit points for the data flowing along the wires. Input-/output pins currently need not be drawn explicitly. This is to keep designing on a whiteboard simple and quick.   Data flowing is of some type, so wires have a type attached to them. And pins have names. If there is only one input pin and output pin on a FU, though, you don´t need to mention them. The default is Process for a single input pin, and Result for a single output pin. But you´re free to give even single pins different names.   There is a shortcut in use to address a certain pin on a destination FU:   The type of the wire is put in parantheses for two reasons. 1. This way a “no-type” wire can be easily denoted, 2. this is a natural way to describe tuples of data.   To describe how much data is flowing, a star can be put next to the wire type:   Nesting – Boards and Parts If more than 5 to 10 FUs need to be put in a flow a FD starts to become hard to understand. To keep diagrams clutter free they can be nested. You can turn any FU into a flow: This leads to Flow-Designs with different levels of abstraction. A in the above illustration is a high level functional unit, A.1 and A.2 are lower level functional units. One of the purposes of Flow-Design is to be able to describe systems on different levels of abstraction and thus make it easier to understand them. Humans use abstraction/decomposition to get a grip on complexity. Flow-Design strives to support this and make levels of abstraction first class citizens for programming. You can read the above illustration like this: Functional units A.1 and A.2 detail what A is supposed to do. The whole of A´s responsibility is decomposed into smaller responsibilities A.1 and A.2. FU A thus does not do anything itself anymore! All A is responsible for is actually accomplished by the collaboration between A.1 and A.2. Since A now is not doing anything anymore except containing A.1 and A.2 functional units are devided into two categories: boards and parts. Boards are just containing other functional units; their sole responsibility is to wire them up. A is a board. Boards thus depend on the functional units nested within them. This dependency is not of a functional nature, though. Boards are not dependent on services provided by nested functional units. They are just concerned with their interface to be able to plug them together. Parts are the workhorses of flows. They contain the real domain logic. They actually transform input into output. However, they do not depend on other functional units. Please note the usage of source and sink in boards. They correspond to input-pins and output-pins of the board.   Implicit Dependencies Nesting functional units leads to a dependency tree. Boards depend on nested functional units, they are the inner nodes of the tree. Parts are independent, they are the leafs: Even though dependencies are the bane of software development, Flow-Design does not usually draw these dependencies. They are implicitly created by visually nesting functional units. And they are harmless. Boards are so simple in their functionality, they are little affected by changes in functional units they are depending on. But functional units are implicitly dependent on more than nested functional units. They are also dependent on the data types of the wires attached to them: This is also natural and thus does not need to be made explicit. And it pertains mainly to parts being dependent. Since boards don´t do anything with regard to a problem domain, they don´t care much about data types. Their infrastructural purpose just needs types of input/output-pins to match.   Explicit Dependencies You could say, Flow-Orientation is about tackling complexity at its root cause: that´s dependencies. “Natural” dependencies are depicted naturally, i.e. implicitly. And whereever possible dependencies are not even created. Functional units don´t know their collaborators within a flow. This is core to Flow-Orientation. That makes for high composability of functional units. A part is as independent of other functional units as a motor is from the rest of the car. And a board is as dependend on nested functional units as a motor is on a spark plug or a crank shaft. With Flow-Design software development moves closer to how hardware is constructed. Implicit dependencies are not enough, though. Sometimes explicit dependencies make designs easier – as counterintuitive this might sound. So FD notation needs a ways to denote explicit dependencies: Data flows along wires. But data does not flow along dependency relations. Instead dependency relations represent service calls. Functional unit C is depending on/calling services on functional unit S. If you want to be more specific, name the services next to the dependency relation: Although you should try to stay clear of explicit dependencies, they are fundamentally ok. See them as a way to add another dimension to a flow. Usually the functionality of the independent FU (“Customer repository” above) is orthogonal to the domain of the flow it is referenced by. If you like emphasize this by using different shapes for dependent and independent FUs like above. Such dependencies can be used to link in resources like databases or shared in-memory state. FUs can not only produce output but also can have side effects. A common pattern for using such explizit dependencies is to hook a GUI into a flow as the source and/or the sink of data: Which can be shortened to: Treat FUs others depend on as boards (with a special non-FD API the dependent part is connected to), but do not embed them in a flow in the diagram they are depended upon.   Attributes of Functional Units Creation and usage of functional units can be modified with attributes. So far the following have shown to be helpful: Singleton: FUs are by default multitons. FUs in the same of different flows with the same name refer to the same functionality, but to different instances. Think of functional units as objects that get instanciated anew whereever they appear in a design. Sometimes though it´s helpful to reuse the same instance of a functional unit; this is always due to valuable state it holds. Signify this by annotating the FU with a “(S)”. Multiton: FUs on which others depend are singletons by default. This is, because they usually are introduced where shared state comes into play. If you want to change them to be a singletons mark them with a “(M)”. Configurable: Some parts need to be configured before the can do they work in a flow. Annotate them with a “(C)” to have them initialized before any data items to be processed by them arrive. Do not assume any order in which FUs are configured. How such configuration is happening is an implementation detail. Entry point: In each design there needs to be a single part where “it all starts”. That´s the entry point for all processing. It´s like Program.Main() in C# programs. Mark the entry point part with an “(E)”. Quite often this will be the GUI part. How the entry point is started is an implementation detail. Just consider it the first FU to start do its job.   Patterns / Standard Parts If more than a single wire is attached to an output-pin that´s called a split (or fork). The same data is flowing on all of the wires. Remember: Flow-Designs are synchronous by default. So a split does not mean data is processed in parallel afterwards. Processing still happens synchronously and thus one branch after another. Do not assume any specific order of the processing on the different branches after the split.   It is common to do a split and let only parts of the original data flow on through the branches. This effectively means a map is needed after a split. This map can be implicit or explicit.   Although FUs can have multiple input-pins it is preferrable in most cases to combine input data from different branches using an explicit join: The default output of a join is a tuple of its input values. The default behavior of a join is to output a value whenever a new input is received. However, to produce its first output a join needs an input for all its input-pins. Other join behaviors can be: reset all inputs after an output only produce output if data arrives on certain input-pins

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  • iPhone SDK vs Windows Phone 7 Series SDK Challenge, Part 1: Hello World!

    In this series, I will be taking sample applications from the iPhone SDK and implementing them on Windows Phone 7 Series.  My goal is to do as much of an apples-to-apples comparison as I can.  This series will be written to not only compare and contrast how easy or difficult it is to complete tasks on either platform, how many lines of code, etc., but Id also like it to be a way for iPhone developers to either get started on Windows Phone 7 Series development, or for developers in general to learn the platform. Heres my methodology: Run the iPhone SDK app in the iPhone Simulator to get a feel for what it does and how it works, without looking at the implementation Implement the equivalent functionality on Windows Phone 7 Series using Silverlight. Compare the two implementations based on complexity, functionality, lines of code, number of files, etc. Add some functionality to the Windows Phone 7 Series app that shows off a way to make the scenario more interesting or leverages an aspect of the platform, or uses a better design pattern to implement the functionality. You can download Microsoft Visual Studio 2010 Express for Windows Phone CTP here, and the Expression Blend 4 Beta here. Hello World! Of course no first post would be allowed if it didnt focus on the hello world scenario.  The iPhone SDK follows that tradition with the Your First iPhone Application walkthrough.  I will say that the developer documentation for iPhone is pretty good.  There are plenty of walkthoughs and they break things down into nicely sized steps and do a good job of bringing the user along.  As expected, this application is quite simple.  It comprises of a text box, a label, and a button.  When you push the button, the label changes to Hello plus the  word you typed into the text box.  Makes perfect sense for a starter application.  Theres not much to this but it covers a few basic elements: Laying out basic UI Handling user input Hooking up events Formatting text     So, lets get started building a similar app for Windows Phone 7 Series! Implementing the UI: UI in Silverlight (and therefore Windows Phone 7) is defined in XAML, which is a declarative XML language also used by WPF on the desktop.  For anyone thats familiar with similar types of markup, its relatively straightforward to learn, but has a lot of power in it once you get it figured out.  Well talk more about that. This UI is very simple.  When I look at this, I note a couple of things: Elements are arranged vertically They are all centered So, lets create our Application and then start with the UI.  Once you have the the VS 2010 Express for Windows Phone tool running, create a new Windows Phone Project, and call it Hello World: Once created, youll see the designer on one side and your XAML on the other: Now, we can create our UI in one of three ways: Use the designer in Visual Studio to drag and drop the components Use the designer in Expression Blend 4 to drag and drop the components Enter the XAML by hand in either of the above Well start with (1), then kind of move to (3) just for instructional value. To develop this UI in the designer: First, delete all of the markup between inside of the Grid element (LayoutRoot).  You should be left with just this XAML for your MainPage.xaml (i shortened all the xmlns declarations below for brevity): 1: <phoneNavigation:PhoneApplicationPage 2: x:Class="HelloWorld.MainPage" 3: xmlns="...[snip]" 4: FontFamily="{StaticResource PhoneFontFamilyNormal}" 5: FontSize="{StaticResource PhoneFontSizeNormal}" 6: Foreground="{StaticResource PhoneForegroundBrush}"> 7:   8: <Grid x:Name="LayoutRoot" Background="{StaticResource PhoneBackgroundBrush}"> 9:   10: </Grid> 11:   12: </phoneNavigation:PhoneApplicationPage> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }   Well be adding XAML at line 9, so thats the important part. Now, Click on the center area of the phone surface Open the Toolbox and double click StackPanel Double click TextBox Double click TextBlock Double click Button That will create the necessary UI elements but they wont be arranged quite right.  Well fix it in a second.    Heres the XAML that we end up with: 1: <StackPanel Height="100" HorizontalAlignment="Left" Margin="10,10,0,0" Name="stackPanel1" VerticalAlignment="Top" Width="200"> 2: <TextBox Height="32" Name="textBox1" Text="TextBox" Width="100" /> 3: <TextBlock Height="23" Name="textBlock1" Text="TextBlock" /> 4: <Button Content="Button" Height="70" Name="button1" Width="160" /> 5: </StackPanel> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The designer does its best at guessing what we want, but in this case we want things to be a bit simpler. So well just clean it up a bit.  We want the items to be centered and we want them to have a little bit of a margin on either side, so heres what we end up with.  Ive also made it match the values and style from the iPhone app: 1: <StackPanel Margin="10"> 2: <TextBox Name="textBox1" HorizontalAlignment="Stretch" Text="You" TextAlignment="Center"/> 3: <TextBlock Name="textBlock1" HorizontalAlignment="Center" Margin="0,100,0,0" Text="Hello You!" /> 4: <Button Name="button1" HorizontalAlignment="Center" Margin="0,150,0,0" Content="Hello"/> 5: </StackPanel> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now lets take a look at what weve done there. Line 1: We removed all of the formatting from the StackPanel, except for Margin, as thats all we need.  Since our parent element is a Grid, by default the StackPanel will be sized to fit in that space.  The Margin says that we want to reserve 10 pixels on each side of the StackPanel. Line 2: Weve set the HorizontalAlignment of the TextBox to Stretch, which says that it should fill its parents size horizontally.  We want to do this so the TextBox is always full-width.  We also set TextAlignment to Center, to center the text. Line 3: In contrast to the TextBox above, we dont care how wide the TextBlock is, just so long as it is big enough for its text.  Thatll happen automatically, so we just set its Horizontal alignment to Center.  We also set a Margin above the TextBlock of 100 pixels to bump it down a bit, per the iPhone UI. Line 4: We do the same things here as in Line 3. Heres how the UI looks in the designer: Believe it or not, were almost done! Implementing the App Logic Now, we want the TextBlock to change its text when the Button is clicked.  In the designer, double click the Button to be taken to the Event Handler for the Buttons Click event.  In that event handler, we take the Text property from the TextBox, and format it into a string, then set it into the TextBlock.  Thats it! 1: private void button1_Click(object sender, RoutedEventArgs e) 2: { 3: string name = textBox1.Text; 4:   5: // if there isn't a name set, just use "World" 6: if (String.IsNullOrEmpty(name)) 7: { 8: name = "World"; 9: } 10:   11: // set the value into the TextBlock 12: textBlock1.Text = String.Format("Hello {0}!", name); 13:   14: } .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; } We use the String.Format() method to handle the formatting for us.    Now all thats left is to test the app in the Windows Phone Emulator and verify it does what we think it does! And it does! Comparing against the iPhone Looking at the iPhone example, there are basically three things that you have to touch as the developer: 1) The UI in the Nib file 2) The app delegate 3) The view controller Counting lines is a bit tricky here, but to try to keep this even, Im going to only count lines of code that I could not have (or would not have) generated with the tooling.  Meaning, Im not counting XAML and Im not counting operations that happen in the Nib file with the XCode designer tool.  So in the case of the above, even though I modified the XAML, I could have done all of those operations using the visual designer tool.  And normally I would have, but the XAML is more instructive (and less steps!).  Im interested in things that I, as the developer have to figure out in code.  Im also not counting lines that just have a curly brace on them, or lines that are generated for me (e.g. method names that are generated for me when I make a connection, etc.) So, by that count, heres what I get from the code listing for the iPhone app found here: HelloWorldAppDelegate.h: 6 HelloWorldAppDelegate.m: 12 MyViewController.h: 8 MyViewController.m: 18 Which gives me a grand total of about 44 lines of code on iPhone.  I really do recommend looking at the iPhone code for a comparison to the above. Now, for the Windows Phone 7 Series application, the only code I typed was in the event handler above Main.Xaml.cs: 4 So a total of 4 lines of code on Windows Phone 7.  And more importantly, the process is just A LOT simpler.  For example, I was surprised that the User Interface Designer in XCode doesnt automatically create instance variables for me and wire them up to the corresponding elements.  I assumed I wouldnt have to write this code myself (and risk getting it wrong!).  I dont need to worry about view controllers or anything.  I just write my code.  This blog post up to this point has covered almost every aspect of this apps development in a few pages.  The iPhone tutorial has 5 top level steps with 2-3 sub sections of each. Now, its worth pointing out that the iPhone development model uses the Model View Controller (MVC) pattern, which is a very flexible and powerful pattern that enforces proper separation of concerns.  But its fairly complex and difficult to understand when you first walk up to it.  Here at Microsoft weve dabbled in MVC a bit, with frameworks like MFC on Visual C++ and with the ASP.NET MVC framework now.  Both are very powerful frameworks.  But one of the reasons weve stayed away from MVC with client UI frameworks is that its difficult to tool.  We havent seen the type of value that beats double click, write code! for the broad set of scenarios. Another thing to think about is how many of those lines of code were focused on my apps functionality?.  Or, the converse of How many lines of code were boilerplate plumbing?  In both examples, the actual number of functional code lines is similar.  I count most of them in MyViewController.m, in the changeGreeting method.  Its about 7 lines of code that do the work of taking the value from the TextBox and putting it into the label.  Versus 4 on the Windows Phone 7 side.  But, unfortunately, on iPhone I still have to write that other 37 lines of code, just to get there. 10% of the code, 1 file instead of 4, its just much simpler. Making Some Tweaks It turns out, I can actually do this application with ZERO  lines of code, if Im willing to change the spec a bit. The data binding functionality in Silverlight is incredibly powerful.  And what I can do is databind the TextBoxs value directly to the TextBlock.  Take some time looking at this XAML below.  Youll see that I have added another nested StackPanel and two more TextBlocks.  Why?  Because thats how I build that string, and the nested StackPanel will lay things out Horizontally for me, as specified by the Orientation property. 1: <StackPanel Margin="10"> 2: <TextBox Name="textBox1" HorizontalAlignment="Stretch" Text="You" TextAlignment="Center"/> 3: <StackPanel Orientation="Horizontal" HorizontalAlignment="Center" Margin="0,100,0,0" > 4: <TextBlock Text="Hello " /> 5: <TextBlock Name="textBlock1" Text="{Binding ElementName=textBox1, Path=Text}" /> 6: <TextBlock Text="!" /> 7: </StackPanel> 8: <Button Name="button1" HorizontalAlignment="Center" Margin="0,150,0,0" Content="Hello" Click="button1_Click" /> 9: </StackPanel> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now, the real action is there in the bolded TextBlock.Text property: Text="{Binding ElementName=textBox1, Path=Text}" .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 does all the heavy lifting.  It sets up a databinding between the TextBox.Text property on textBox1 and the TextBlock.Text property on textBlock1. As I change the text of the TextBox, the label updates automatically. In fact, I dont even need the button any more, so I could get rid of that altogether.  And no button means no event handler.  No event handler means no C# code at all.  Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • Quick Quips on QR Codes

    - by Tim Dexter
    Yes, I'm an alliterating all-star; I missed my calling as a newspaper headline writer. I have recently received questions from several folks on support for QR codes. You know them they are everywhere you look, even here! How does Publisher handle QR codes then? In theory, exactly the same way we handle any other 2D barcode font. We need the font file, a mapping entry and an encoding class. With those three pieces we can embed QR codes into any output. To test the theory, I went off to IDAutomation, I have worked with them and many customers over the years and their fonts and encoders have worked great and have been very reliable. They kindly provide demo fonts which has made my life so much easier to be able to write posts like this. Their QR font and encoder is a little tough to find. I started here and then hit the Demo Now button. On the next page I hit the right hand Demo Now button. In the resulting zip file you'll need two files: AdditionalFonts.zip >> Automation2DFonts >> TrueType >> IDAutomation2D.ttf Java Class Encoder >> IDAutomation_JavaFontEncoder_QRCode.jar - the QRBarcodeExample.java is useful to see how to call the encoder. The font file needs to be installed into the windows/fonts directory, just copy and paste it in using file explorer and windows will install it for you. Remember, we are using the demo font here and you'll see if you get your phones decoder to looks a the font above there is a fixed string 'DEMO' at the beginning. You want that removed? Go buy the font from the IDAutomation folks. The Encoder Next you need to create your encoding wrapper class. Publisher does ship a class but its compiled and I do not recommend trying to modify it, you can just build your own. I have loaded up my class here. You do not need to be a java guru, its pretty straightforward. I'd recommend a java IDE like JDeveloper from a convenience point of view. I have annotated my class and added a main method to it so you can test your encoders from JDeveloper without having to deploy them first. You can load up the project form the zip file straight into JDeveloper.Next, take a look at IDAutomation's example java class and you'll see: QRCodeEncoder qre=new QRCodeEncoder();  String DataToEncode = "IDAutmation Inc.";  boolean ApplyTilde = false;  int EncodingMode = 0;  int Version = 0;  int ErrorCorrectionLevel = 0;  System.out.println( qre.FontEncode(DataToEncode, ApplyTilde, EncodingMode, Version, ErrorCorrectionLevel) ); You'll need to check what settings you need to set for the ApplyTilde, EncodingMode, Version and ErrorCorrectionLevel. They are covered in the user guide from IDAutomation here. If you do not want to hard code the values in the encoder then you can quite easily externalize them and read the values from a text file. I have not covered that scenario here, I'm going with IDAutomation's defaults and my phone app is reading the fonts no problem. Now you know how to call the encoder, you need to incorporate it into your encoder wrapper class. From my sample class:       Class[] clazz = new Class[] { "".getClass() };        ENCODERS.put("code128a",mUtility.getClass().getMethod("code128a", clazz));       ENCODERS.put("code128b",mUtility.getClass().getMethod("code128b", clazz));       ENCODERS.put("code128c",mUtility.getClass().getMethod("code128c", clazz));       ENCODERS.put("qrcode",mUtility.getClass().getMethod("qrcode", clazz)); I just added a new entry to register the encoder method 'qrcode' (in red). Then I created a new method inside the class to call the IDAutomation encoder. /** Call to IDAutomations QR Code encoder. Passing the data to encode      Returning the encoded string to the template for formatting **/ public static final String qrcode (String DataToEncode) {   QRCodeEncoder qre=new QRCodeEncoder();    boolean ApplyTilde = false;    int EncodingMode = 0;    int Version = 0;    int ErrorCorrectionLevel = 0; return qre.FontEncode(DataToEncode, ApplyTilde, EncodingMode, Version, ErrorCorrectionLevel); } Almost the exact same code in their sample class. The DataToEncode string is passed in rather than hardcoded of course. With the class done you can now compile it, but you need to ensure that the IDAutomation_JavaFontEncoder_QRCode.jar is in the classpath. In JDeveloper, open the project properties >> Libraries and Classpaths and then add the jar to the list. You'll need the publisher jars too. You can find those in the jlib directory in your Template Builder for Word directory.Note! In my class, I have used package oracle.psbi.barcode; As my package spec, yours will be different but you need to note it for later. Once you have it compiling without errors you will need to generate a jar file to keep it in. In JDeveloper highlight your project node >> New >> Deployment Profile >> JAR file. Once you have created the descriptor, just take the defaults. It will tell you where the jar is located. Go get it and then its time to copy it and the IDAutomation jar into the Template Builder for Word directory structure. Deploying the jars On your windows machine locate the jlib directory under the Template Builder for Word install directory. On my machine its here, F:\Program Files\Oracle\BI Publisher\BI Publisher Desktop\Template Builder for Word\jlib. Copy both of the jar files into the directory. The next step is to get the jars into the classpath for the Word plugin so that Publisher can find your wrapper class and it can then find the IDAutomation encoder. The most consistent way I have found so far, is to open up the RTF2PDF.jar in the same directory and make some mods. First make a backup of the jar file then open it using winzip or 7zip or similar and get into the META-INF directory. In there is a file, MANIFEST.MF. This contains the classpath for the plugin, open it in an editor and add the jars to the end of the classpath list. In mine I have: Manifest-Version: 1.0 Class-Path: ./activation.jar ./mail.jar ./xdochartstyles.jar ./bicmn.jar ./jewt4.jar ./share.jar ./bipres.jar ./xdoparser.jar ./xdocore.jar ./xmlparserv2.jar ./xmlparserv2-904.jar  ./i18nAPI_v3.jar ./versioninfo.jar ./barcodejar.jar ./IDAutomation_JavaFontEncoder_QRCode.jar Main-Class: RTF2PDF I have put in carriage returns above to make the Class-Path: entry more readable, make sure yours is all on one line. Be sure to use the ./ as a prefix to the jar name. Ensure the file is saved inside the jar file 7zip and winzip both have popups asking if you want to update the file in the jar file.Now you have the jars on the classpath, the Publisher plugin will be able to find our classes at run time. Referencing the Font The next step is to reference the font location so that the rendering engine can find it and embed a subset into the PDF output. Remember the other output formats rely on the font being present on the machine that is opening the document. The PDF is the only truly portable format. Inside the config directory under the Template Builder for Word install directory, mine is here, F:\Program Files\Oracle\BI Publisher\BI Publisher Desktop\Template Builder for Word\config. You'll find the file, 'xdo example.cfg'. Rename it to xdo.cfg and open it in a text editor. In the fonts section, create a new entry:       <font family="IDAutomation2D" style="normal" weight="normal">              <truetype path="C:\windows\fonts\IDAutomation2D.ttf" />       </font> Note, 'IDAutomation2D' (in red) is the same name as you can see when you open MSWord and look for the QRCode font. This must match exactly. When Publisher looks at the fonts in the RTF template at runtime it will see 'IDAutomation2D' it will then look at its font mapping entries to find where that font file resides on the disk. If the names do not match or the font is not present then the font will not get used and it will fall back on Helvetica. Building the Template Now you have the data encoder and the font in place and mapped; you can use it in the template. The two commands you will need to have present are: <?register-barcode-vendor:'ENCODER WRAPPER CLASS'; 'ENCODER NAME'?> for my encoder I have: <?register-barcode-vendor:'oracle.psbi.barcode.BarcodeUtil'; 'MyBarcodeEncoder'?> Notice the two parameters for the command. The first provides the package 'path' and class name (remember I said you need to remember that above.)The second is the name of the encoder, in my case 'MyBarcodeEncoder'. Check my full encoder class in the zip linked below to see where I named it. You can change it to something else, no problem.This command needs to be near the top of the template. The second command is the encoding command: <?format-barcode:DATAT_TO_ENCODE;'ENCODER_METHOD_NAME';'ENCODER_NAME'?> for my command I have <?format-barcode:DATATEXT;'qrcode';'MyBarcodeEncoder'?>DATATEXT is the XML element that contains the text to be encoded. If you want to hard code a piece of text just surround it with single quotes. qrcode is the name of my encoder method that calls the IDAutomation encoder. Remember this.MyBarcodeEncoder is the name of my encoder. Repetition? Yes but its needed again. Both of these commands are put inside their own form fields. Do not apply the QRCode font to the second field just yet. Lets make sure the encoder is working. Run you template with some data and you should get something like this for your encoded data: AHEEEHAPPJOPMOFADIPFJKDCLPAHEEEHA BNFFFNBPJGMDIDJPFOJGIGBLMPBNFFFNB APIBOHFJCFBNKHGGBMPFJFJLJBKGOMNII OANKPJFFLEPLDNPCLMNGNIJIHFDNLJFEH FPLFLHFHFILKFBLOIGMDFCFLGJGOPJJME CPIACDFJPBGDODOJCHALJOBPECKMOEDDF MFFNFNEPKKKCHAIHCHPCFFLDAHFHAGLMK APBBBPAPLDKNKJKKGIPDLKGMGHDDEPHLN HHHHHHHPHPHHPHPPHPPPPHHPHHPHPHPHP Grooovy huh? If you do not get the encoded text then go back and check that your jars are in the right spot and that you have the MANIFEST.MF file updated correctly. Once you do get the encoded text, highlight the field and apply the IDAutomation2D font to it. Then re-run the report and you will hopefully see the QR code in your output. If not, go back and check the xdo.cfg entry and make sure its in the right place and the font location is correct. That's it, you now have QR codes in Publisher outputs. Everything I have written above, has been tested with the 5.6.3, 10.1.3.4.2 codelines. I'll be testing the 11g code in the next day or two and will update you with any changes. One thing I have not covered yet and will do in the next few days is how to deploy all of this to your server. Look out for a follow up post. One note on the apparent white lines in the font (see the image above). Once printed they disappear and even viewing the code on a screen with the white lines, my phone app is still able to read and interpret the contents no problem. I have zipped up my encoder wrapper class as a JDeveloper 11.1.1.6 project here. Just dig into the src directories to find the BarcodeUtil.java file if you just want the code. I have put comments into the file to hopefully help the novice java programmer out. Happy QR'ing!

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  • How to setup linux permissions the WWW folder?

    - by Xeoncross
    Updated Summery The /var/www directory is owned by root:root which means that no one can use it and it's entirely useless. Since we all want a web server that actually works (and no-one should be logging in as "root"), then we need to fix this. Only two entities need access. PHP/Perl/Ruby/Python all need access to the folders and files since they create many of them (i.e. /uploads/). These scripting languages should be running under nginx or apache (or even some other thing like FastCGI for PHP). The developers How do they get access? I know that someone, somewhere has done this before. With however-many billions of websites out there you would think that there would be more information on this topic. I know that 777 is full read/write/execute permission for owner/group/other. So this doesn't seem to be needed as it leaves random users full permissions. What permissions are need to be used on /var/www so that... Source control like git or svn Users in a group like "websites" (or even added to "www-data") Servers like apache or lighthttpd And PHP/Perl/Ruby can all read, create, and run files (and directories) there? If I'm correct, Ruby and PHP scripts are not "executed" directly - but passed to an interpreter. So there is no need for execute permission on files in /var/www...? Therefore, it seems like the correct permission would be chmod -R 1660 which would make all files shareable by these four entities all files non-executable by mistake block everyone else from the directory entirely set the permission mode to "sticky" for all future files Is this correct? Update: I just realized that files and directories might need different permissions - I was talking about files above so i'm not sure what the directory permissions would need to be. Update 2: The folder structure of /var/www changes drastically as one of the four entities above are always adding (and sometimes removing) folders and sub folders many levels deep. They also create and remove files that the other 3 entities might need read/write access to. Therefore, the permissions need to do the four things above for both files and directories. Since non of them should need execute permission (see question about ruby/php above) I would assume that rw-rw-r-- permission would be all that is needed and completely safe since these four entities are run by trusted personal (see #2) and all other users on the system only have read access. Update 3: This is for personal development machines and private company servers. No random "web customers" like a shared host. Update 4: This article by slicehost seems to be the best at explaining what is needed to setup permissions for your www folder. However, I'm not sure what user or group apache/nginx with PHP OR svn/git run as and how to change them. Update 5: I have (I think) finally found a way to get this all to work (answer below). However, I don't know if this is the correct and SECURE way to do this. Therefore I have started a bounty. The person that has the best method of securing and managing the www directory wins.

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  • Diving into OpenStack Network Architecture - Part 2 - Basic Use Cases

    - by Ronen Kofman
      rkofman Normal rkofman 4 138 2014-06-05T03:38:00Z 2014-06-05T05:04:00Z 3 2735 15596 Oracle Corporation 129 36 18295 12.00 Clean Clean false false false false EN-US X-NONE HE /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:Arial; mso-bidi-theme-font:minor-bidi; mso-bidi-language:AR-SA;} In the previous post we reviewed several network components including Open vSwitch, Network Namespaces, Linux Bridges and veth pairs. In this post we will take three simple use cases and see how those basic components come together to create a complete SDN solution in OpenStack. With those three use cases we will review almost the entire network setup and see how all the pieces work together. The use cases we will use are: 1.       Create network – what happens when we create network and how can we create multiple isolated networks 2.       Launch a VM – once we have networks we can launch VMs and connect them to networks. 3.       DHCP request from a VM – OpenStack can automatically assign IP addresses to VMs. This is done through local DHCP service controlled by OpenStack Neutron. We will see how this service runs and how does a DHCP request and response look like. In this post we will show connectivity, we will see how packets get from point A to point B. We first focus on how a configured deployment looks like and only later we will discuss how and when the configuration is created. Personally I found it very valuable to see the actual interfaces and how they connect to each other through examples and hands on experiments. After the end game is clear and we know how the connectivity works, in a later post, we will take a step back and explain how Neutron configures the components to be able to provide such connectivity.  We are going to get pretty technical shortly and I recommend trying these examples on your own deployment or using the Oracle OpenStack Tech Preview. Understanding these three use cases thoroughly and how to look at them will be very helpful when trying to debug a deployment in case something does not work. Use case #1: Create Network Create network is a simple operation it can be performed from the GUI or command line. When we create a network in OpenStack the network is only available to the tenant who created it or it could be defined as “shared” and then it can be used by all tenants. A network can have multiple subnets but for this demonstration purpose and for simplicity we will assume that each network has exactly one subnet. Creating a network from the command line will look like this: # neutron net-create net1 Created a new network: +---------------------------+--------------------------------------+ | Field                     | Value                                | +---------------------------+--------------------------------------+ | admin_state_up            | True                                 | | id                        | 5f833617-6179-4797-b7c0-7d420d84040c | | name                      | net1                                 | | provider:network_type     | vlan                                 | | provider:physical_network | default                              | | provider:segmentation_id  | 1000                                 | | shared                    | False                                | | status                    | ACTIVE                               | | subnets                   |                                      | | tenant_id                 | 9796e5145ee546508939cd49ad59d51f     | +---------------------------+--------------------------------------+ Creating a subnet for this network will look like this: # neutron subnet-create net1 10.10.10.0/24 Created a new subnet: +------------------+------------------------------------------------+ | Field            | Value                                          | +------------------+------------------------------------------------+ | allocation_pools | {"start": "10.10.10.2", "end": "10.10.10.254"} | | cidr             | 10.10.10.0/24                                  | | dns_nameservers  |                                                | | enable_dhcp      | True                                           | | gateway_ip       | 10.10.10.1                                     | | host_routes      |                                                | | id               | 2d7a0a58-0674-439a-ad23-d6471aaae9bc           | | ip_version       | 4                                              | | name             |                                                | | network_id       | 5f833617-6179-4797-b7c0-7d420d84040c           | | tenant_id        | 9796e5145ee546508939cd49ad59d51f               | +------------------+------------------------------------------------+ We now have a network and a subnet, on the network topology view this looks like this: Now let’s dive in and see what happened under the hood. Looking at the control node we will discover that a new namespace was created: # ip netns list qdhcp-5f833617-6179-4797-b7c0-7d420d84040c   The name of the namespace is qdhcp-<network id> (see above), let’s look into the namespace and see what’s in it: # ip netns exec qdhcp-5f833617-6179-4797-b7c0-7d420d84040c ip addr 1: lo: <LOOPBACK,UP,LOWER_UP> mtu 65536 qdisc noqueue state UNKNOWN     link/loopback 00:00:00:00:00:00 brd 00:00:00:00:00:00     inet 127.0.0.1/8 scope host lo     inet6 ::1/128 scope host        valid_lft forever preferred_lft forever 12: tap26c9b807-7c: <BROADCAST,UP,LOWER_UP> mtu 1500 qdisc noqueue state UNKNOWN     link/ether fa:16:3e:1d:5c:81 brd ff:ff:ff:ff:ff:ff     inet 10.10.10.3/24 brd 10.10.10.255 scope global tap26c9b807-7c     inet6 fe80::f816:3eff:fe1d:5c81/64 scope link        valid_lft forever preferred_lft forever   We see two interfaces in the namespace, one is the loopback and the other one is an interface called “tap26c9b807-7c”. This interface has the IP address of 10.10.10.3 and it will also serve dhcp requests in a way we will see later. Let’s trace the connectivity of the “tap26c9b807-7c” interface from the namespace.  First stop is OVS, we see that the interface connects to bridge  “br-int” on OVS: # ovs-vsctl show 8a069c7c-ea05-4375-93e2-b9fc9e4b3ca1     Bridge "br-eth2"         Port "br-eth2"             Interface "br-eth2"                 type: internal         Port "eth2"             Interface "eth2"         Port "phy-br-eth2"             Interface "phy-br-eth2"     Bridge br-ex         Port br-ex             Interface br-ex                 type: internal     Bridge br-int         Port "int-br-eth2"             Interface "int-br-eth2"         Port "tap26c9b807-7c"             tag: 1             Interface "tap26c9b807-7c"                 type: internal         Port br-int             Interface br-int                 type: internal     ovs_version: "1.11.0"   In the picture above we have a veth pair which has two ends called “int-br-eth2” and "phy-br-eth2", this veth pair is used to connect two bridge in OVS "br-eth2" and "br-int". In the previous post we explained how to check the veth connectivity using the ethtool command. It shows that the two are indeed a pair: # ethtool -S int-br-eth2 NIC statistics:      peer_ifindex: 10 . .   #ip link . . 10: phy-br-eth2: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qdisc pfifo_fast state UP qlen 1000 . . Note that “phy-br-eth2” is connected to a bridge called "br-eth2" and one of this bridge's interfaces is the physical link eth2. This means that the network which we have just created has created a namespace which is connected to the physical interface eth2. eth2 is the “VM network” the physical interface where all the virtual machines connect to where all the VMs are connected. About network isolation: OpenStack supports creation of multiple isolated networks and can use several mechanisms to isolate the networks from one another. The isolation mechanism can be VLANs, VxLANs or GRE tunnels, this is configured as part of the initial setup in our deployment we use VLANs. When using VLAN tagging as an isolation mechanism a VLAN tag is allocated by Neutron from a pre-defined VLAN tags pool and assigned to the newly created network. By provisioning VLAN tags to the networks Neutron allows creation of multiple isolated networks on the same physical link.  The big difference between this and other platforms is that the user does not have to deal with allocating and managing VLANs to networks. The VLAN allocation and provisioning is handled by Neutron which keeps track of the VLAN tags, and responsible for allocating and reclaiming VLAN tags. In the example above net1 has the VLAN tag 1000, this means that whenever a VM is created and connected to this network the packets from that VM will have to be tagged with VLAN tag 1000 to go on this particular network. This is true for namespace as well, if we would like to connect a namespace to a particular network we have to make sure that the packets to and from the namespace are correctly tagged when they reach the VM network. In the example above we see that the namespace interface “tap26c9b807-7c” has vlan tag 1 assigned to it, if we examine OVS we see that it has flows which modify VLAN tag 1 to VLAN tag 1000 when a packet goes to the VM network on eth2 and vice versa. We can see this using the dump-flows command on OVS for packets going to the VM network we see the modification done on br-eth2: #  ovs-ofctl dump-flows br-eth2 NXST_FLOW reply (xid=0x4):  cookie=0x0, duration=18669.401s, table=0, n_packets=857, n_bytes=163350, idle_age=25, priority=4,in_port=2,dl_vlan=1 actions=mod_vlan_vid:1000,NORMAL  cookie=0x0, duration=165108.226s, table=0, n_packets=14, n_bytes=1000, idle_age=5343, hard_age=65534, priority=2,in_port=2 actions=drop  cookie=0x0, duration=165109.813s, table=0, n_packets=1671, n_bytes=213304, idle_age=25, hard_age=65534, priority=1 actions=NORMAL   For packets coming from the interface to the namespace we see the following modification: #  ovs-ofctl dump-flows br-int NXST_FLOW reply (xid=0x4):  cookie=0x0, duration=18690.876s, table=0, n_packets=1610, n_bytes=210752, idle_age=1, priority=3,in_port=1,dl_vlan=1000 actions=mod_vlan_vid:1,NORMAL  cookie=0x0, duration=165130.01s, table=0, n_packets=75, n_bytes=3686, idle_age=4212, hard_age=65534, priority=2,in_port=1 actions=drop  cookie=0x0, duration=165131.96s, table=0, n_packets=863, n_bytes=160727, idle_age=1, hard_age=65534, priority=1 actions=NORMAL   To summarize we can see that when a user creates a network Neutron creates a namespace and this namespace is connected through OVS to the “VM network”. OVS also takes care of tagging the packets from the namespace to the VM network with the correct VLAN tag and knows to modify the VLAN for packets coming from VM network to the namespace. Now let’s see what happens when a VM is launched and how it is connected to the “VM network”. Use case #2: Launch a VM Launching a VM can be done from Horizon or from the command line this is how we do it from Horizon: Attach the network: And Launch Once the virtual machine is up and running we can see the associated IP using the nova list command : # nova list +--------------------------------------+--------------+--------+------------+-------------+-----------------+ | ID                                   | Name         | Status | Task State | Power State | Networks        | +--------------------------------------+--------------+--------+------------+-------------+-----------------+ | 3707ac87-4f5d-4349-b7ed-3a673f55e5e1 | Oracle Linux | ACTIVE | None       | Running     | net1=10.10.10.2 | +--------------------------------------+--------------+--------+------------+-------------+-----------------+ The nova list command shows us that the VM is running and that the IP 10.10.10.2 is assigned to this VM. Let’s trace the connectivity from the VM to VM network on eth2 starting with the VM definition file. The configuration files of the VM including the virtual disk(s), in case of ephemeral storage, are stored on the compute node at/var/lib/nova/instances/<instance-id>/. Looking into the VM definition file ,libvirt.xml,  we see that the VM is connected to an interface called “tap53903a95-82” which is connected to a Linux bridge called “qbr53903a95-82”: <interface type="bridge">       <mac address="fa:16:3e:fe:c7:87"/>       <source bridge="qbr53903a95-82"/>       <target dev="tap53903a95-82"/>     </interface>   Looking at the bridge using the brctl show command we see this: # brctl show bridge name     bridge id               STP enabled     interfaces qbr53903a95-82          8000.7e7f3282b836       no              qvb53903a95-82                                                         tap53903a95-82    The bridge has two interfaces, one connected to the VM (“tap53903a95-82 “) and another one ( “qvb53903a95-82”) connected to “br-int” bridge on OVS: # ovs-vsctl show 83c42f80-77e9-46c8-8560-7697d76de51c     Bridge "br-eth2"         Port "br-eth2"             Interface "br-eth2"                 type: internal         Port "eth2"             Interface "eth2"         Port "phy-br-eth2"             Interface "phy-br-eth2"     Bridge br-int         Port br-int             Interface br-int                 type: internal         Port "int-br-eth2"             Interface "int-br-eth2"         Port "qvo53903a95-82"             tag: 3             Interface "qvo53903a95-82"     ovs_version: "1.11.0"   As we showed earlier “br-int” is connected to “br-eth2” on OVS using the veth pair int-br-eth2,phy-br-eth2 and br-eth2 is connected to the physical interface eth2. The whole flow end to end looks like this: VM è tap53903a95-82 (virtual interface)è qbr53903a95-82 (Linux bridge) è qvb53903a95-82 (interface connected from Linux bridge to OVS bridge br-int) è int-br-eth2 (veth one end) è phy-br-eth2 (veth the other end) è eth2 physical interface. The purpose of the Linux Bridge connecting to the VM is to allow security group enforcement with iptables. Security groups are enforced at the edge point which are the interface of the VM, since iptables nnot be applied to OVS bridges we use Linux bridge to apply them. In the future we hope to see this Linux Bridge going away rules.  VLAN tags: As we discussed in the first use case net1 is using VLAN tag 1000, looking at OVS above we see that qvo41f1ebcf-7c is tagged with VLAN tag 3. The modification from VLAN tag 3 to 1000 as we go to the physical network is done by OVS  as part of the packet flow of br-eth2 in the same way we showed before. To summarize, when a VM is launched it is connected to the VM network through a chain of elements as described here. During the packet from VM to the network and back the VLAN tag is modified. Use case #3: Serving a DHCP request coming from the virtual machine In the previous use cases we have shown that both the namespace called dhcp-<some id> and the VM end up connecting to the physical interface eth2  on their respective nodes, both will tag their packets with VLAN tag 1000.We saw that the namespace has an interface with IP of 10.10.10.3. Since the VM and the namespace are connected to each other and have interfaces on the same subnet they can ping each other, in this picture we see a ping from the VM which was assigned 10.10.10.2 to the namespace: The fact that they are connected and can ping each other can become very handy when something doesn’t work right and we need to isolate the problem. In such case knowing that we should be able to ping from the VM to the namespace and back can be used to trace the disconnect using tcpdump or other monitoring tools. To serve DHCP requests coming from VMs on the network Neutron uses a Linux tool called “dnsmasq”,this is a lightweight DNS and DHCP service you can read more about it here. If we look at the dnsmasq on the control node with the ps command we see this: dnsmasq --no-hosts --no-resolv --strict-order --bind-interfaces --interface=tap26c9b807-7c --except-interface=lo --pid-file=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/pid --dhcp-hostsfile=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/host --dhcp-optsfile=/var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/opts --leasefile-ro --dhcp-range=tag0,10.10.10.0,static,120s --dhcp-lease-max=256 --conf-file= --domain=openstacklocal The service connects to the tap interface in the namespace (“--interface=tap26c9b807-7c”), If we look at the hosts file we see this: # cat  /var/lib/neutron/dhcp/5f833617-6179-4797-b7c0-7d420d84040c/host fa:16:3e:fe:c7:87,host-10-10-10-2.openstacklocal,10.10.10.2   If you look at the console output above you can see the MAC address fa:16:3e:fe:c7:87 which is the VM MAC. This MAC address is mapped to IP 10.10.10.2 and so when a DHCP request comes with this MAC dnsmasq will return the 10.10.10.2.If we look into the namespace at the time we initiate a DHCP request from the VM (this can be done by simply restarting the network service in the VM) we see the following: # ip netns exec qdhcp-5f833617-6179-4797-b7c0-7d420d84040c tcpdump -n 19:27:12.191280 IP 0.0.0.0.bootpc > 255.255.255.255.bootps: BOOTP/DHCP, Request from fa:16:3e:fe:c7:87, length 310 19:27:12.191666 IP 10.10.10.3.bootps > 10.10.10.2.bootpc: BOOTP/DHCP, Reply, length 325   To summarize, the DHCP service is handled by dnsmasq which is configured by Neutron to listen to the interface in the DHCP namespace. Neutron also configures dnsmasq with the combination of MAC and IP so when a DHCP request comes along it will receive the assigned IP. Summary In this post we relied on the components described in the previous post and saw how network connectivity is achieved using three simple use cases. These use cases gave a good view of the entire network stack and helped understand how an end to end connection is being made between a VM on a compute node and the DHCP namespace on the control node. One conclusion we can draw from what we saw here is that if we launch a VM and it is able to perform a DHCP request and receive a correct IP then there is reason to believe that the network is working as expected. We saw that a packet has to travel through a long list of components before reaching its destination and if it has done so successfully this means that many components are functioning properly. In the next post we will look at some more sophisticated services Neutron supports and see how they work. We will see that while there are some more components involved for the most part the concepts are the same. @RonenKofman

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  • C#/.NET Little Wonders: Interlocked CompareExchange()

    - by James Michael Hare
    Once again, in this series of posts I look at the parts of the .NET Framework that may seem trivial, but can help improve your code by making it easier to write and maintain. The index of all my past little wonders posts can be found here. Two posts ago, I discussed the Interlocked Add(), Increment(), and Decrement() methods (here) for adding and subtracting values in a thread-safe, lightweight manner.  Then, last post I talked about the Interlocked Read() and Exchange() methods (here) for safely and efficiently reading and setting 32 or 64 bit values (or references).  This week, we’ll round out the discussion by talking about the Interlocked CompareExchange() method and how it can be put to use to exchange a value if the current value is what you expected it to be. Dirty reads can lead to bad results Many of the uses of Interlocked that we’ve explored so far have centered around either reading, setting, or adding values.  But what happens if you want to do something more complex such as setting a value based on the previous value in some manner? Perhaps you were creating an application that reads a current balance, applies a deposit, and then saves the new modified balance, where of course you’d want that to happen atomically.  If you read the balance, then go to save the new balance and between that time the previous balance has already changed, you’ll have an issue!  Think about it, if we read the current balance as $400, and we are applying a new deposit of $50.75, but meanwhile someone else deposits $200 and sets the total to $600, but then we write a total of $450.75 we’ve lost $200! Now, certainly for int and long values we can use Interlocked.Add() to handles these cases, and it works well for that.  But what if we want to work with doubles, for example?  Let’s say we wanted to add the numbers from 0 to 99,999 in parallel.  We could do this by spawning several parallel tasks to continuously add to a total: 1: double total = 0; 2:  3: Parallel.For(0, 10000, next => 4: { 5: total += next; 6: }); Were this run on one thread using a standard for loop, we’d expect an answer of 4,999,950,000 (the sum of all numbers from 0 to 99,999).  But when we run this in parallel as written above, we’ll likely get something far off.  The result of one of my runs, for example, was 1,281,880,740.  That is way off!  If this were banking software we’d be in big trouble with our clients.  So what happened?  The += operator is not atomic, it will read in the current value, add the result, then store it back into the total.  At any point in all of this another thread could read a “dirty” current total and accidentally “skip” our add.   So, to clean this up, we could use a lock to guarantee concurrency: 1: double total = 0.0; 2: object locker = new object(); 3:  4: Parallel.For(0, count, next => 5: { 6: lock (locker) 7: { 8: total += next; 9: } 10: }); Which will give us the correct result of 4,999,950,000.  One thing to note is that locking can be heavy, especially if the operation being locked over is trivial, or the life of the lock is a high percentage of the work being performed concurrently.  In the case above, the lock consumes pretty much all of the time of each parallel task – and the task being locked on is relatively trivial. Now, let me put in a disclaimer here before we go further: For most uses, lock is more than sufficient for your needs, and is often the simplest solution!    So, if lock is sufficient for most needs, why would we ever consider another solution?  The problem with locking is that it can suspend execution of your thread while it waits for the signal that the lock is free.  Moreover, if the operation being locked over is trivial, the lock can add a very high level of overhead.  This is why things like Interlocked.Increment() perform so well, instead of locking just to perform an increment, we perform the increment with an atomic, lockless method. As with all things performance related, it’s important to profile before jumping to the conclusion that you should optimize everything in your path.  If your profiling shows that locking is causing a high level of waiting in your application, then it’s time to consider lighter alternatives such as Interlocked. CompareExchange() – Exchange existing value if equal some value So let’s look at how we could use CompareExchange() to solve our problem aboveThe general syntax of CompareExchange() is: T CompareExchange<T>(ref T location, T newValue, T expectedValue) If the value in location == expectedValue, then newValue is exchanged.  Either way, the value in location (before exchange) is returned. Actually, CompareExchange() is not one method, but a family of overloaded methods that can take int, long, float, double, pointers, or references.  It cannot take other value types (that is, can’t CompareExchange() two DateTime instances directly).  Also keep in mind that the version that takes any reference type (the generic overload) only checks for reference equality, it does not call any overridden Equals(). So how does this help us?  Well, we can grab the current total, and exchange the new value if total hasn’t changed.  This would look like this: 1: // grab the snapshot 2: double current = total; 3:  4: // if the total hasn’t changed since I grabbed the snapshot, then 5: // set it to the new total 6: Interlocked.CompareExchange(ref total, current + next, current); So what the code above says is: if the amount in total (1st arg) is the same as the amount in current (3rd arg), then set total to current + next (2nd arg).  This check and exchange pair is atomic (and thus thread-safe). This works if total is the same as our snapshot in current, but the problem, is what happens if they aren’t the same?  Well, we know that in either case we will get the previous value of total (before the exchange), back as a result.  Thus, we can test this against our snapshot to see if it was the value we expected: 1: // if the value returned is != current, then our snapshot must be out of date 2: // which means we didn't (and shouldn't) apply current + next 3: if (Interlocked.CompareExchange(ref total, current + next, current) != current) 4: { 5: // ooops, total was not equal to our snapshot in current, what should we do??? 6: } So what do we do if we fail?  That’s up to you and the problem you are trying to solve.  It’s possible you would decide to abort the whole transaction, or perhaps do a lightweight spin and try again.  Let’s try that: 1: double current = total; 2:  3: // make first attempt... 4: if (Interlocked.CompareExchange(ref total, current + i, current) != current) 5: { 6: // if we fail, go into a spin wait, spin, and try again until succeed 7: var spinner = new SpinWait(); 8:  9: do 10: { 11: spinner.SpinOnce(); 12: current = total; 13: } 14: while (Interlocked.CompareExchange(ref total, current + i, current) != current); 15: } 16:  This is not trivial code, but it illustrates a possible use of CompareExchange().  What we are doing is first checking to see if we succeed on the first try, and if so great!  If not, we create a SpinWait and then repeat the process of SpinOnce(), grab a fresh snapshot, and repeat until CompareExchnage() succeeds.  You may wonder why not a simple do-while here, and the reason it’s more efficient to only create the SpinWait until we absolutely know we need one, for optimal efficiency. Though not as simple (or maintainable) as a simple lock, this will perform better in many situations.  Comparing an unlocked (and wrong) version, a version using lock, and the Interlocked of the code, we get the following average times for multiple iterations of adding the sum of 100,000 numbers: 1: Unlocked money average time: 2.1 ms 2: Locked money average time: 5.1 ms 3: Interlocked money average time: 3 ms So the Interlocked.CompareExchange(), while heavier to code, came in lighter than the lock, offering a good compromise of safety and performance when we need to reduce contention. CompareExchange() - it’s not just for adding stuff… So that was one simple use of CompareExchange() in the context of adding double values -- which meant we couldn’t have used the simpler Interlocked.Add() -- but it has other uses as well. If you think about it, this really works anytime you want to create something new based on a current value without using a full lock.  For example, you could use it to create a simple lazy instantiation implementation.  In this case, we want to set the lazy instance only if the previous value was null: 1: public static class Lazy<T> where T : class, new() 2: { 3: private static T _instance; 4:  5: public static T Instance 6: { 7: get 8: { 9: // if current is null, we need to create new instance 10: if (_instance == null) 11: { 12: // attempt create, it will only set if previous was null 13: Interlocked.CompareExchange(ref _instance, new T(), (T)null); 14: } 15:  16: return _instance; 17: } 18: } 19: } So, if _instance == null, this will create a new T() and attempt to exchange it with _instance.  If _instance is not null, then it does nothing and we discard the new T() we created. This is a way to create lazy instances of a type where we are more concerned about locking overhead than creating an accidental duplicate which is not used.  In fact, the BCL implementation of Lazy<T> offers a similar thread-safety choice for Publication thread safety, where it will not guarantee only one instance was created, but it will guarantee that all readers get the same instance.  Another possible use would be in concurrent collections.  Let’s say, for example, that you are creating your own brand new super stack that uses a linked list paradigm and is “lock free”.  We could use Interlocked.CompareExchange() to be able to do a lockless Push() which could be more efficient in multi-threaded applications where several threads are pushing and popping on the stack concurrently. Yes, there are already concurrent collections in the BCL (in .NET 4.0 as part of the TPL), but it’s a fun exercise!  So let’s assume we have a node like this: 1: public sealed class Node<T> 2: { 3: // the data for this node 4: public T Data { get; set; } 5:  6: // the link to the next instance 7: internal Node<T> Next { get; set; } 8: } Then, perhaps, our stack’s Push() operation might look something like: 1: public sealed class SuperStack<T> 2: { 3: private volatile T _head; 4:  5: public void Push(T value) 6: { 7: var newNode = new Node<int> { Data = value, Next = _head }; 8:  9: if (Interlocked.CompareExchange(ref _head, newNode, newNode.Next) != newNode.Next) 10: { 11: var spinner = new SpinWait(); 12:  13: do 14: { 15: spinner.SpinOnce(); 16: newNode.Next = _head; 17: } 18: while (Interlocked.CompareExchange(ref _head, newNode, newNode.Next) != newNode.Next); 19: } 20: } 21:  22: // ... 23: } Notice a similar paradigm here as with adding our doubles before.  What we are doing is creating the new Node with the data to push, and with a Next value being the original node referenced by _head.  This will create our stack behavior (LIFO – Last In, First Out).  Now, we have to set _head to now refer to the newNode, but we must first make sure it hasn’t changed! So we check to see if _head has the same value we saved in our snapshot as newNode.Next, and if so, we set _head to newNode.  This is all done atomically, and the result is _head’s original value, as long as the original value was what we assumed it was with newNode.Next, then we are good and we set it without a lock!  If not, we SpinWait and try again. Once again, this is much lighter than locking in highly parallelized code with lots of contention.  If I compare the method above with a similar class using lock, I get the following results for pushing 100,000 items: 1: Locked SuperStack average time: 6 ms 2: Interlocked SuperStack average time: 4.5 ms So, once again, we can get more efficient than a lock, though there is the cost of added code complexity.  Fortunately for you, most of the concurrent collection you’d ever need are already created for you in the System.Collections.Concurrent (here) namespace – for more information, see my Little Wonders – The Concurent Collections Part 1 (here), Part 2 (here), and Part 3 (here). Summary We’ve seen before how the Interlocked class can be used to safely and efficiently add, increment, decrement, read, and exchange values in a multi-threaded environment.  In addition to these, Interlocked CompareExchange() can be used to perform more complex logic without the need of a lock when lock contention is a concern. The added efficiency, though, comes at the cost of more complex code.  As such, the standard lock is often sufficient for most thread-safety needs.  But if profiling indicates you spend a lot of time waiting for locks, or if you just need a lock for something simple such as an increment, decrement, read, exchange, etc., then consider using the Interlocked class’s methods to reduce wait. Technorati Tags: C#,CSharp,.NET,Little Wonders,Interlocked,CompareExchange,threading,concurrency

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  • Symantec Protection Suite and System Recovery 2011 Desktop Edition

    - by rihatum
    I am re-posting this as my previous question was being treated as if I am "Shopping or seeking Product Recommendations" even though I was NOT - BTW they have deleted my comments too which were not offensive in nature. anyway - I have re-phrased some parts of my question and I hope SF Admins "Do Not Modify / Edit" this one - will be most grateful for that. I have a lot of respect for the People who visit this SITE and help others ! Just To clarify : Just to go by SF rules - I am not seeking someone to Design this solution, I am simply seeking real world examples, experiences, technical expert opinions / suggestions, any tips or tricks they may have or any problems they may have faced while doing something similar above with these products. I am also not asking for Capacity Planning for Storage, We have done some research and I am seeking Expert Assurance / Suggestions. We (our company) are planning to deploy Symantec Endpoint Protection and Symantec Desktop Recovery 2011 Desktop Edition to our 3000 - 4000 workstations (Windows7 32 and 64) with a few 100s with Windows XP 32/64 Bit. I have read the implementation guide for SEP and have read tech-notes for Desktop Recovery 2011. Our team have planned to deploy this as follows : 1 x dedicated SQL 2008R2 for Symantec Endpoint Protection (Instead of using the Embedded Database) 1 x Dedicated SQL 2008R2 for Symantec Desktop Recovery 2011 (Instead of using the Embedded Database) 1 x Dedicated W2K8 R2 Box for the SEPM (Symantec Endpoint Protection Manager - Mgmt. APP) 1 x Dedicated W2K8 R2 Box for the Symantec Desktop Recovery 2011 Management Application Agent Deployment : As per Symantec Documentation for both of the above, an agent can be pushed via the Mgmt. Application (provided no firewalls are blocking ports required etc. - we have Windows firewall disabled already). Server Hardware : Per SQL Server : 16GB RAM + SAS DISKS + Dual XEON, RAID-10 for the SQL DB or I can always mount a LUN from our existing Hitachi or EMC SAN. SEPM Server : 16GB RAM + SAS DISKS + DUAL XEON System Recovery MGMT SERVER : 16GB RAM + SAS DISKS + DUAL XEON Above is the initial plan we have for 3000 - 4000 client workstation (Windows) Now my Questions :-) a) If we had these users distributed amongst two sites with AD DC / GC in each site, How would I restrict SEPM and Desktop Mgmt. solution to only check for users in their respective site ? b) At present all users are under one building but we are going to move some dept. to a new location (with dedicated connectivity), How would we control which SEPM / MGMT Server is responsible for which site ? c) We have netbackup in our environment backing up other servers, I am planning to protect these 4 (2 x SQL, 1 x SEPM, 1 x System Recovery Mgmt. Server) via netbackup or I can use System recovery 2011 server edition on all 4 of these boxes as well. (License is not an issue as we have the complete symantec portfolio included in our license). d) Now - Saving Desktop backups - What strategies have you implemented ? Any best practice recommendation for a large user base ? I was thinking to either mount a LUN from our Hitachi SAN on the Symantec Recovery Server itself or backup to the users hard drive locally and then copy it over to a network location ? Suggestions welcome :-) If you have anything to add / correct - that will be really helpful before diving into the actual implementation phase. Will be most grateful with your suggestions, recommendations and corrections with above - Many Thanks !

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  • How to Load Oracle Tables From Hadoop Tutorial (Part 5 - Leveraging Parallelism in OSCH)

    - by Bob Hanckel
    Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 Using OSCH: Beyond Hello World In the previous post we discussed a “Hello World” example for OSCH focusing on the mechanics of getting a toy end-to-end example working. In this post we are going to talk about how to make it work for big data loads. We will explain how to optimize an OSCH external table for load, paying particular attention to Oracle’s DOP (degree of parallelism), the number of external table location files we use, and the number of HDFS files that make up the payload. We will provide some rules that serve as best practices when using OSCH. The assumption is that you have read the previous post and have some end to end OSCH external tables working and now you want to ramp up the size of the loads. Using OSCH External Tables for Access and Loading OSCH external tables are no different from any other Oracle external tables.  They can be used to access HDFS content using Oracle SQL: SELECT * FROM my_hdfs_external_table; or use the same SQL access to load a table in Oracle. INSERT INTO my_oracle_table SELECT * FROM my_hdfs_external_table; To speed up the load time, you will want to control the degree of parallelism (i.e. DOP) and add two SQL hints. ALTER SESSION FORCE PARALLEL DML PARALLEL  8; ALTER SESSION FORCE PARALLEL QUERY PARALLEL 8; INSERT /*+ append pq_distribute(my_oracle_table, none) */ INTO my_oracle_table SELECT * FROM my_hdfs_external_table; There are various ways of either hinting at what level of DOP you want to use.  The ALTER SESSION statements above force the issue assuming you (the user of the session) are allowed to assert the DOP (more on that in the next section).  Alternatively you could embed additional parallel hints directly into the INSERT and SELECT clause respectively. /*+ parallel(my_oracle_table,8) *//*+ parallel(my_hdfs_external_table,8) */ Note that the "append" hint lets you load a target table by reserving space above a given "high watermark" in storage and uses Direct Path load.  In other doesn't try to fill blocks that are already allocated and partially filled. It uses unallocated blocks.  It is an optimized way of loading a table without incurring the typical resource overhead associated with run-of-the-mill inserts.  The "pq_distribute" hint in this context unifies the INSERT and SELECT operators to make data flow during a load more efficient. Finally your target Oracle table should be defined with "NOLOGGING" and "PARALLEL" attributes.   The combination of the "NOLOGGING" and use of the "append" hint disables REDO logging, and its overhead.  The "PARALLEL" clause tells Oracle to try to use parallel execution when operating on the target table. Determine Your DOP It might feel natural to build your datasets in Hadoop, then afterwards figure out how to tune the OSCH external table definition, but you should start backwards. You should focus on Oracle database, specifically the DOP you want to use when loading (or accessing) HDFS content using external tables. The DOP in Oracle controls how many PQ slaves are launched in parallel when executing an external table. Typically the DOP is something you want to Oracle to control transparently, but for loading content from Hadoop with OSCH, it's something that you will want to control. Oracle computes the maximum DOP that can be used by an Oracle user. The maximum value that can be assigned is an integer value typically equal to the number of CPUs on your Oracle instances, times the number of cores per CPU, times the number of Oracle instances. For example, suppose you have a RAC environment with 2 Oracle instances. And suppose that each system has 2 CPUs with 32 cores. The maximum DOP would be 128 (i.e. 2*2*32). In point of fact if you are running on a production system, the maximum DOP you are allowed to use will be restricted by the Oracle DBA. This is because using a system maximum DOP can subsume all system resources on Oracle and starve anything else that is executing. Obviously on a production system where resources need to be shared 24x7, this can’t be allowed to happen. The use cases for being able to run OSCH with a maximum DOP are when you have exclusive access to all the resources on an Oracle system. This can be in situations when your are first seeding tables in a new Oracle database, or there is a time where normal activity in the production database can be safely taken off-line for a few hours to free up resources for a big incremental load. Using OSCH on high end machines (specifically Oracle Exadata and Oracle BDA cabled with Infiniband), this mode of operation can load up to 15TB per hour. The bottom line is that you should first figure out what DOP you will be allowed to run with by talking to the DBAs who manage the production system. You then use that number to derive the number of location files, and (optionally) the number of HDFS data files that you want to generate, assuming that is flexible. Rule 1: Find out the maximum DOP you will be allowed to use with OSCH on the target Oracle system Determining the Number of Location Files Let’s assume that the DBA told you that your maximum DOP was 8. You want the number of location files in your external table to be big enough to utilize all 8 PQ slaves, and you want them to represent equally balanced workloads. Remember location files in OSCH are metadata lists of HDFS files and are created using OSCH’s External Table tool. They also represent the workload size given to an individual Oracle PQ slave (i.e. a PQ slave is given one location file to process at a time, and only it will process the contents of the location file.) Rule 2: The size of the workload of a single location file (and the PQ slave that processes it) is the sum of the content size of the HDFS files it lists For example, if a location file lists 5 HDFS files which are each 100GB in size, the workload size for that location file is 500GB. The number of location files that you generate is something you control by providing a number as input to OSCH’s External Table tool. Rule 3: The number of location files chosen should be a small multiple of the DOP Each location file represents one workload for one PQ slave. So the goal is to keep all slaves busy and try to give them equivalent workloads. Obviously if you run with a DOP of 8 but have 5 location files, only five PQ slaves will have something to do and the other three will have nothing to do and will quietly exit. If you run with 9 location files, then the PQ slaves will pick up the first 8 location files, and assuming they have equal work loads, will finish up about the same time. But the first PQ slave to finish its job will then be rescheduled to process the ninth location file, potentially doubling the end to end processing time. So for this DOP using 8, 16, or 32 location files would be a good idea. Determining the Number of HDFS Files Let’s start with the next rule and then explain it: Rule 4: The number of HDFS files should try to be a multiple of the number of location files and try to be relatively the same size In our running example, the DOP is 8. This means that the number of location files should be a small multiple of 8. Remember that each location file represents a list of unique HDFS files to load, and that the sum of the files listed in each location file is a workload for one Oracle PQ slave. The OSCH External Table tool will look in an HDFS directory for a set of HDFS files to load.  It will generate N number of location files (where N is the value you gave to the tool). It will then try to divvy up the HDFS files and do its best to make sure the workload across location files is as balanced as possible. (The tool uses a greedy algorithm that grabs the biggest HDFS file and delegates it to a particular location file. It then looks for the next biggest file and puts in some other location file, and so on). The tools ability to balance is reduced if HDFS file sizes are grossly out of balance or are too few. For example suppose my DOP is 8 and the number of location files is 8. Suppose I have only 8 HDFS files, where one file is 900GB and the others are 100GB. When the tool tries to balance the load it will be forced to put the singleton 900GB into one location file, and put each of the 100GB files in the 7 remaining location files. The load balance skew is 9 to 1. One PQ slave will be working overtime, while the slacker PQ slaves are off enjoying happy hour. If however the total payload (1600 GB) were broken up into smaller HDFS files, the OSCH External Table tool would have an easier time generating a list where each workload for each location file is relatively the same.  Applying Rule 4 above to our DOP of 8, we could divide the workload into160 files that were approximately 10 GB in size.  For this scenario the OSCH External Table tool would populate each location file with 20 HDFS file references, and all location files would have similar workloads (approximately 200GB per location file.) As a rule, when the OSCH External Table tool has to deal with more and smaller files it will be able to create more balanced loads. How small should HDFS files get? Not so small that the HDFS open and close file overhead starts having a substantial impact. For our performance test system (Exadata/BDA with Infiniband), I compared three OSCH loads of 1 TiB. One load had 128 HDFS files living in 64 location files where each HDFS file was about 8GB. I then did the same load with 12800 files where each HDFS file was about 80MB size. The end to end load time was virtually the same. However when I got ridiculously small (i.e. 128000 files at about 8MB per file), it started to make an impact and slow down the load time. What happens if you break rules 3 or 4 above? Nothing draconian, everything will still function. You just won’t be taking full advantage of the generous DOP that was allocated to you by your friendly DBA. The key point of the rules articulated above is this: if you know that HDFS content is ultimately going to be loaded into Oracle using OSCH, it makes sense to chop them up into the right number of files roughly the same size, derived from the DOP that you expect to use for loading. Next Steps So far we have talked about OLH and OSCH as alternative models for loading. That’s not quite the whole story. They can be used together in a way that provides for more efficient OSCH loads and allows one to be more flexible about scheduling on a Hadoop cluster and an Oracle Database to perform load operations. The next lesson will talk about Oracle Data Pump files generated by OLH, and loaded using OSCH. It will also outline the pros and cons of using various load methods.  This will be followed up with a final tutorial lesson focusing on how to optimize OLH and OSCH for use on Oracle's engineered systems: specifically Exadata and the BDA. /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin-top:0in; mso-para-margin-right:0in; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0in; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin;}

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  • What's the lowest cost, legal, Microsoft server stack you can assemble?

    - by McKAMEY
    Assuming that you have an app infrastructure that generally only requires: ASP.NET MVC / C# / .NET Database or NoSQL data store (must be accessible from C#) Here's the challenge to you server gods: What is the least expensive configuration that will allow you to deploy to production in a way that doesn't break any licensing rules? In what ways does this solution differ from the "standard" Microsoft deployment scenario? Where does this solution's performance break down once the app begins to scale? I'm not concerned about the hardware, only the server software itself. I would love to hear about any solutions you've personally put into production. Especially if they are unique alternatives. For ideas, consider some of the possible variations, a) any Microsoft server solutions where they have lowered the barrier to entry to compete with OSS, or b) any OSS alternatives to Microsoft products which perform at a similar level. An example of a): SQL Server 2008 Express Edition SP1 is a 100% free version of SQL Server which will scale to the needs of many smaller / early stage applications. An example of b): running the Mono Framework on Linux. An example of differing from the "standard" stack: running Mono on Linux will require a completely different server OS familiarity. None of the Windows-based knowledge really transfers. An example of breaking down under scale: SQL Server Express will only scale to 1GB of memory and 4GB of disk storage. After that point, the application will need to move to one of the paid versions of SQL Server.

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  • How do I prevent TCP connection freezes over an OpenVPN network?

    - by Jason R
    New details added at the end of this question; it's possible that I'm zeroing in on the cause. I have a UDP OpenVPN-based VPN set up in tap mode (I need tap because I need the VPN to pass multicast packets, which doesn't seem to be possible with tun networks) with a handful of clients across the Internet. I've been experiencing frequent TCP connection freezes over the VPN. That is, I will establish a TCP connection (e.g. an SSH connection, but other protocols have similar issues), and at some point during the session, it seems that traffic will cease being transmitted over that TCP session. This seems to be related to points at which large data transfers occur, such as if I execute an ls command in an SSH session, or if I cat a long log file. Some Google searches turn up a number of answers like this previous one on Server Fault, indicating that the likely culprit is an MTU issue: that during periods of high traffic, the VPN is trying to send packets that get dropped somewhere in the pipes between the VPN endpoints. The above-linked answer suggests using the following OpenVPN configuration settings to mitigate the problem: fragment 1400 mssfix This should limit the MTU used on the VPN to 1400 bytes and fix the TCP maximum segment size to prevent the generation of any packets larger than that. This seems to mitigate the problem a bit, but I still frequently see the freezes. I've tried a number of sizes as arguments to the fragment directive: 1200, 1000, 576, all with similar results. I can't think of any strange network topology between the two ends that could trigger such a problem: the VPN server is running on a pfSense machine connected directly to the Internet, and my client is also connected directly to the Internet at another location. One other strange piece of the puzzle: if I run the tracepath utility, then that seems to band-aid the problem. A sample run looks like: [~]$ tracepath -n 192.168.100.91 1: 192.168.100.90 0.039ms pmtu 1500 1: 192.168.100.91 40.823ms reached 1: 192.168.100.91 19.846ms reached Resume: pmtu 1500 hops 1 back 64 The above run is between two clients on the VPN: I initiated the trace from 192.168.100.90 to the destination of 192.168.100.91. Both clients were configured with fragment 1200; mssfix; in an attempt to limit the MTU used on the link. The above results would seem to suggest that tracepath was able to detect a path MTU of 1500 bytes between the two clients. I would assume that it would be somewhat smaller due to the fragmentation settings specified in the OpenVPN configuration. I found that result somewhat strange. Even stranger, however: if I have a TCP connection in the stalled state (e.g. an SSH session with a directory listing that froze in the middle), then executing the tracepath command shown above causes the connection to start up again! I can't figure out any reasonable explanation for why this would be the case, but I feel like this might be pointing toward a solution to ultimately eradicate the problem. Does anyone have any recommendations for other things to try? Edit: I've come back and looked at this a bit further, and have found only more confounding information: I set the OpenVPN connection to fragment at 1400 bytes, as shown above. Then, I connected to the VPN from across the Internet and used Wireshark to look at the UDP packets that were sent to the VPN server while the stall occurred. None were greater than the specified 1400 byte count, so the fragmentation seems to be functioning properly. To verify that even a 1400-byte MTU would be sufficient, I pinged the VPN server using the following (Linux) command: ping <host> -s 1450 -M do This (I believe) sends a 1450-byte packet with fragmentation disabled (I at least verified that it didn't work if I set it to an obviously-too-large value like 1600 bytes). These seem to work just fine; I get replies back from the host with no issue. So, maybe this isn't an MTU issue at all. I'm just confused as to what else it might be! Edit 2: The rabbit hole just keeps getting deeper: I've now isolated the problem a bit more. It seems to be related to the exact OS that the VPN client uses. I have successfully duplicated the problem on at least three Ubuntu machines (versions 12.04 through 13.04). I can reliably duplicate an SSH connection freeze within a minute or so by just cat-ing a large log file. However, if I do the same test using a CentOS 6 machine as a client, then I don't see the problem! I've tested using the exact same OpenVPN client version as I was using on the Ubuntu machines. I can cat log files for hours without seeing the connection freeze. This seems to provide some insight as to the ultimate cause, but I'm just not sure what that insight is. I have examined the traffic over the VPN using Wireshark. I'm not a TCP expert, so I'm not sure what to make of the gory details, but the gist is that at some point, a UDP packet gets dropped due to the limited bandwidth of the Internet link, causing TCP retransmissions inside the VPN tunnel. On the CentOS client, these retransmissions occur properly and things move on happily. At some point with the Ubuntu clients, though, the remote end starts retransmitting the same TCP segment over and over (with the transmit delay increasing between each retransmission). The client sends what looks like a valid TCP ACK to each retransmission, but the remote end still continues to transmit the same TCP segment periodically. This extends ad infinitum and the connection stalls. My question here would be: Does anyone have any recommendations for how to troubleshoot and/or determine the root cause of the TCP issue? It's as if the remote end isn't accepting the ACK messages sent by the VPN client. One common difference between the CentOS node and the various Ubuntu releases is that Ubuntu has a much more recent Linux kernel version (from 3.2 in Ubuntu 12.04 to 3.8 in 13.04). A pointer to some new kernel bug maybe? I'm assuming that if that were so, then I wouldn't be the only one experiencing the problem; I don't think this seems like a particularly exotic setup.

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  • Missing Memory on Windows Server 2008

    - by Chris Lively
    I have a windows 2008 x64 server with 8GB of RAM installed. Task Manager and Resource Monitor both insist that 7.5GB of the RAM is in use. However, the memory list under Processes (Memory Private Bytes) doesn't add up. I do have Show Processes from all users checked and hand adding the numbers I come up with about 3.5GB of RAM. I also looked at the latest copy of SysInternals Process Explorer. And neither the Private Bytes or Working Set adds up to more than about 3.5GB of RAM in use. What's going on? ===== Update: I bounced the server to see what would happen with the memory utilization. After boot and regular operations began it sat at 3GB of RAM usage. 18 hours later, it's back up to 6.8GB of usage with no indication as to where the additional 3.5GB or so of RAM is being used. Here are links to screen shots of the resource monitor and task manager: Resource Monitor Task Manager Update 2: Well, I believe I located the problem. When I detached one of the larger databases from my sql server the amount of ram shown as "in use" dropped drastically. The Memory Private Bytes count barely moved. So I'm guessing that SQL server has some way of allocating memory where it doesn't really show up in any of the monitors. I went further and created a new database file, then transferred all of the data from the one I detached. Even though it has the same data, and the same transactions going through it, the memory in use has stayed low. Maybe there was some corruption in the DB? I'll leave it to the DB gods and go searching for another "problem" ;)

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  • Help with Design for Vacation Tracking System (C#/.NET/Access/WebServices/SOA/Excel) [closed]

    - by Aaronaught
    I have been tasked with developing a system for tracking our company's paid time-off (vacation, sick days, etc.) At the moment we are using an Excel spreadsheet on a shared network drive, and it works pretty well, but we are concerned that we won't be able to "trust" employees forever and sometimes we run into locking issues when two people try to open the spreadsheet at once. So we are trying to build something a little more robust. I would like some input on this design in terms of maintainability, scalability, extensibility, etc. It's a pretty simple workflow we need to represent right now: I started with a basic MS Access schema like this: Employees (EmpID int, EmpName varchar(50), AllowedDays int) Vacations (VacationID int, EmpID int, BeginDate datetime, EndDate datetime) But we don't want to spend a lot of time building a schema and database like this and have to change it later, so I think I am going to go with something that will be easier to expand through configuration. Right now the vacation table has this schema: Vacations (VacationID int, PropName varchar(50), PropValue varchar(50)) And the table will be populated with data like this: VacationID | PropName | PropValue -----------+--------------+------------------ 1 | EmpID | 4 1 | EmpName | James Jones 1 | Reason | Vacation 1 | BeginDate | 2/24/2010 1 | EndDate | 2/30/2010 1 | Destination | Spectate Swamp 2 | ... | ... I think this is a pretty good, extensible design, we can easily add new properties to the vacation like the destination or maybe approval status, etc. I wasn't too sure how to go about managing the database of valid properties, I thought of putting them in a separate PropNames table but it gets complicated to manage all the different data types and people say that you shouldn't put CLR type names into a SQL database, so I decided to use XML instead, here is the schema: <VacationProperties> <PropertyNames>EmpID,EmpName,Reason,BeginDate,EndDate,Destination</PropertyNames> <PropertyTypes>System.Int32,System.String,System.String,System.DateTime,System.DateTime,System.String</PropertyTypes> <PropertiesRequired>true,true,false,true,true,false</PropertiesRequired> </VacationProperties> I might need more fields than that, I'm not completely sure. I'm parsing the XML like this (would like some feedback on the parsing code): string xml = File.ReadAllText("properties.xml"); Match m = Regex.Match(xml, "<(PropertyNames)>(.*?)</PropertyNames>"; string[] pn = m.Value.Split(','); // do the same for PropertyTypes, PropertiesRequired Then I use the following code to persist configuration changes to the database: string sql = "DROP TABLE VacationProperties"; sql = sql + " CREATE TABLE VacationProperties "; sql = sql + "(PropertyName varchar(100), PropertyType varchar(100) "; sql = sql + "IsRequired varchar(100))"; for (int i = 0; i < pn.Length; i++) { sql = sql + " INSERT VacationProperties VALUES (" + pn[i] + "," + pt[i] + "," + pv[i] + ")"; } // GlobalConnection is a singleton new SqlCommand(sql, GlobalConnection.Instance).ExecuteReader(); So far so good, but after a few days of this I then realized that a lot of this was just a more specific kind of a generic workflow which could be further abstracted, and instead of writing all of this boilerplate plumbing code I could just come up with a workflow and plug it into a workflow engine like Windows Workflow Foundation and have the users configure it: In order to support routing these configurations throw the workflow system, it seemed natural to implement generic XML Web Services for this instead of just using an XML file as above. I've used this code to implement the Web Services: public class VacationConfigurationService : WebService { [WebMethod] public void UpdateConfiguration(string xml) { // Above code goes here } } Which was pretty easy, although I'm still working on a way to validate that XML against some kind of schema as there's no error-checking yet. I also created a few different services for other operations like VacationSubmissionService, VacationReportService, VacationDataService, VacationAuthenticationService, etc. The whole Service Oriented Architecture looks like this: And because the workflow itself might change, I have been working on a way to integrate the WF workflow system with MS Visio, which everybody at the office already knows how to use so they could make changes pretty easily. We have a diagram that looks like the following (it's kind of hard to read but the main items are Activities, Authenticators, Validators, Transformers, Processors, and Data Connections, they're all analogous to the services in the SOA diagram above). The requirements for this system are: (Note - I don't control these, they were given to me by management) Main workflow must interface with Excel spreadsheet, probably through VBA macros (to ease the transition to the new system) Alerts should integrate with MS Outlook, Lotus Notes, and SMS (text messages). We also want to interface it with the company Voice Mail system but that is not a "hard" requirement. Performance requirements: Must handle 250,000 Transactions Per Second Should be able to handle up to 20,000 employees (right now we have 3) 99.99% uptime ("four nines") expected Must be secure against outside hacking, but users cannot be required to enter a username/password. Platforms: Must support Windows XP/Vista/7, Linux, iPhone, Blackberry, DOS 2.0, VAX, IRIX, PDP-11, Apple IIc. Time to complete: 6 to 8 weeks. My questions are: Is this a good design for the system so far? Am I using all of the recommended best practices for these technologies? How do I integrate the Visio diagram above with the Windows Workflow Foundation to call the ConfigurationService and persist workflow changes? Am I missing any important components? Will this be extensible enough to support any scenario via end-user configuration? Will the system scale to the above performance requirements? Will we need any expensive hardware to run it? Are there any "gotchas" I should know about with respect to cross-platform compatibility? For example would it be difficult to convert this to an iPhone app? How long would you expect this to take? (We've dedicated 1 week for testing so I'm thinking maybe 5 weeks?) Many thanks for your advices, Aaron

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  • Session memory – who’s this guy named Max and what’s he doing with my memory?

    - by extended_events
    SQL Server MVP Jonathan Kehayias (blog) emailed me a question last week when he noticed that the total memory used by the buffers for an event session was larger than the value he specified for the MAX_MEMORY option in the CREATE EVENT SESSION DDL. The answer here seems like an excellent subject for me to kick-off my new “401 – Internals” tag that identifies posts where I pull back the curtains a bit and let you peek into what’s going on inside the extended events engine. In a previous post (Option Trading: Getting the most out of the event session options) I explained that we use a set of buffers to store the event data before  we write the event data to asynchronous targets. The MAX_MEMORY along with the MEMORY_PARTITION_MODE defines how big each buffer will be. Theoretically, that means that I can predict the size of each buffer using the following formula: max memory / # of buffers = buffer size If it was that simple I wouldn’t be writing this post. I’ll take “boundary” for 64K Alex For a number of reasons that are beyond the scope of this blog, we create event buffers in 64K chunks. The result of this is that the buffer size indicated by the formula above is rounded up to the next 64K boundary and that is the size used to create the buffers. If you think visually, this means that the graph of your max_memory option compared to the actual buffer size that results will look like a set of stairs rather than a smooth line. You can see this behavior by looking at the output of dm_xe_sessions, specifically the fields related to the buffer sizes, over a range of different memory inputs: Note: This test was run on a 2 core machine using per_cpu partitioning which results in 5 buffers. (Seem my previous post referenced above for the math behind buffer count.) input_memory_kb total_regular_buffers regular_buffer_size total_buffer_size 637 5 130867 654335 638 5 130867 654335 639 5 130867 654335 640 5 196403 982015 641 5 196403 982015 642 5 196403 982015 This is just a segment of the results that shows one of the “jumps” between the buffer boundary at 639 KB and 640 KB. You can verify the size boundary by doing the math on the regular_buffer_size field, which is returned in bytes: 196403 – 130867 = 65536 bytes 65536 / 1024 = 64 KB The relationship between the input for max_memory and when the regular_buffer_size is going to jump from one 64K boundary to the next is going to change based on the number of buffers being created. The number of buffers is dependent on the partition mode you choose. If you choose any partition mode other than NONE, the number of buffers will depend on your hardware configuration. (Again, see the earlier post referenced above.) With the default partition mode of none, you always get three buffers, regardless of machine configuration, so I generated a “range table” for max_memory settings between 1 KB and 4096 KB as an example. start_memory_range_kb end_memory_range_kb total_regular_buffers regular_buffer_size total_buffer_size 1 191 NULL NULL NULL 192 383 3 130867 392601 384 575 3 196403 589209 576 767 3 261939 785817 768 959 3 327475 982425 960 1151 3 393011 1179033 1152 1343 3 458547 1375641 1344 1535 3 524083 1572249 1536 1727 3 589619 1768857 1728 1919 3 655155 1965465 1920 2111 3 720691 2162073 2112 2303 3 786227 2358681 2304 2495 3 851763 2555289 2496 2687 3 917299 2751897 2688 2879 3 982835 2948505 2880 3071 3 1048371 3145113 3072 3263 3 1113907 3341721 3264 3455 3 1179443 3538329 3456 3647 3 1244979 3734937 3648 3839 3 1310515 3931545 3840 4031 3 1376051 4128153 4032 4096 3 1441587 4324761 As you can see, there are 21 “steps” within this range and max_memory values below 192 KB fall below the 64K per buffer limit so they generate an error when you attempt to specify them. Max approximates True as memory approaches 64K The upshot of this is that the max_memory option does not imply a contract for the maximum memory that will be used for the session buffers (Those of you who read Take it to the Max (and beyond) know that max_memory is really only referring to the event session buffer memory.) but is more of an estimate of total buffer size to the nearest higher multiple of 64K times the number of buffers you have. The maximum delta between your initial max_memory setting and the true total buffer size occurs right after you break through a 64K boundary, for example if you set max_memory = 576 KB (see the green line in the table), your actual buffer size will be closer to 767 KB in a non-partitioned event session. You get “stepped up” for every 191 KB block of initial max_memory which isn’t likely to cause a problem for most machines. Things get more interesting when you consider a partitioned event session on a computer that has a large number of logical CPUs or NUMA nodes. Since each buffer gets “stepped up” when you break a boundary, the delta can get much larger because it’s multiplied by the number of buffers. For example, a machine with 64 logical CPUs will have 160 buffers using per_cpu partitioning or if you have 8 NUMA nodes configured on that machine you would have 24 buffers when using per_node. If you’ve just broken through a 64K boundary and get “stepped up” to the next buffer size you’ll end up with total buffer size approximately 10240 KB and 1536 KB respectively (64K * # of buffers) larger than max_memory value you might think you’re getting. Using per_cpu partitioning on large machine has the most impact because of the large number of buffers created. If the amount of memory being used by your system within these ranges is important to you then this is something worth paying attention to and considering when you configure your event sessions. The DMV dm_xe_sessions is the tool to use to identify the exact buffer size for your sessions. In addition to the regular buffers (read: event session buffers) you’ll also see the details for large buffers if you have configured MAX_EVENT_SIZE. The “buffer steps” for any given hardware configuration should be static within each partition mode so if you want to have a handy reference available when you configure your event sessions you can use the following code to generate a range table similar to the one above that is applicable for your specific machine and chosen partition mode. DECLARE @buf_size_output table (input_memory_kb bigint, total_regular_buffers bigint, regular_buffer_size bigint, total_buffer_size bigint) DECLARE @buf_size int, @part_mode varchar(8) SET @buf_size = 1 -- Set to the begining of your max_memory range (KB) SET @part_mode = 'per_cpu' -- Set to the partition mode for the table you want to generate WHILE @buf_size <= 4096 -- Set to the end of your max_memory range (KB) BEGIN     BEGIN TRY         IF EXISTS (SELECT * from sys.server_event_sessions WHERE name = 'buffer_size_test')             DROP EVENT SESSION buffer_size_test ON SERVER         DECLARE @session nvarchar(max)         SET @session = 'create event session buffer_size_test on server                         add event sql_statement_completed                         add target ring_buffer                         with (max_memory = ' + CAST(@buf_size as nvarchar(4)) + ' KB, memory_partition_mode = ' + @part_mode + ')'         EXEC sp_executesql @session         SET @session = 'alter event session buffer_size_test on server                         state = start'         EXEC sp_executesql @session         INSERT @buf_size_output (input_memory_kb, total_regular_buffers, regular_buffer_size, total_buffer_size)             SELECT @buf_size, total_regular_buffers, regular_buffer_size, total_buffer_size FROM sys.dm_xe_sessions WHERE name = 'buffer_size_test'     END TRY     BEGIN CATCH         INSERT @buf_size_output (input_memory_kb)             SELECT @buf_size     END CATCH     SET @buf_size = @buf_size + 1 END DROP EVENT SESSION buffer_size_test ON SERVER SELECT MIN(input_memory_kb) start_memory_range_kb, MAX(input_memory_kb) end_memory_range_kb, total_regular_buffers, regular_buffer_size, total_buffer_size from @buf_size_output group by total_regular_buffers, regular_buffer_size, total_buffer_size Thanks to Jonathan for an interesting question and a chance to explore some of the details of Extended Event internals. - Mike

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  • How to pass XML to DB using XMLTYPE

    - by James Taylor
    Probably not a common use case but I have seen it pop up from time to time. The question how do I pass XML from a queue or web service and insert it into a DB table using XMLTYPE.In this example I create a basic table with the field PAYLOAD of type XMLTYPE. I then take the full XML payload of the web service and insert it into that database for auditing purposes.I use SOA Suite 11.1.1.2 using composite and mediator to link the web service with the DB adapter.1. Insert Database Objects Normal 0 false false false MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} --Create XML_EXAMPLE_TBL Normal 0 false false false MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} CREATE TABLE XML_EXAMPLE_TBL (PAYLOAD XMLTYPE); Normal 0 false false false MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} --Create procedure LOAD_TEST_XML Normal 0 false false false MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} CREATE or REPLACE PROCEDURE load_test_xml (xmlFile in CLOB) IS   BEGIN     INSERT INTO xml_example_tbl (payload) VALUES (XMLTYPE(xmlFile));   --Handle the exceptions EXCEPTION   WHEN OTHERS THEN     raise_application_error(-20101, 'Exception occurred in loadPurchaseOrder procedure :'||SQLERRM || ' **** ' || xmlFile ); END load_test_xml; / 2. Creating New SOA Project TestXMLTYPE in JDeveloperIn JDeveloper either create a new Application or open an existing Application you want to put this work.Under File -> New -> SOA Tier -> SOA Project   Provide a name for the Project, e.g. TestXMLType Choose Empty Composite When selected Empty Composite click Finish.3. Create Database Connection to Stored ProcedureA Blank composite will be displayed. From the Component Palette drag a Database Adapter to the  External References panel. and configure the Database Adapter Wizard to connect to the DB procedure created above.Provide a service name InsertXML Select a Database connection where you installed the table and procedure above. If it doesn't exist create a new one. Select Call a Stored Procedure or Function then click NextChoose the schema you installed your Procedure in step 1 and query for the LOAD_TEST_XML procedure.Click Next for the remaining screens until you get to the end, then click Finish to complete the database adapter wizard.4. Create the Web Service InterfaceDownload this sample schema that will be used as the input for the web service. It does not matter what schema you use this solution will work with any. Feel free to use your own if required. singleString.xsd Drag from the component palette the Web Service to the Exposed Services panel on the component.Provide a name InvokeXMLLoad for the service, and click the cog icon.Click the magnify glass for the URL to browse to the location where you downloaded the xml schema above.  Import the schema file by selecting the import schema iconBrowse to the location to where you downloaded the singleString.xsd above.Click OK for the Import Schema File, then select the singleString node of the imported schema.Accept all the defaults until you get back to the Web Service wizard screen. The click OK. This step has created a WSDL based on the schema we downloaded earlier.Your composite should now look something like this now.5. Create the Mediator Routing Rules Drag a Mediator component into the middle of the Composite called ComponentsGive the name of Route, and accept the defaultsLink the services up to the Mediator by connecting the reference points so your Composite looks like this.6. Perform Translations between Web Service and the Database Adapter.From the Composite double click the Route Mediator to show the Map Plan. Select the transformation icon to create the XSLT translation file.Choose Create New Mapper File and accept the defaults.From the Component Palette drag the get-content-as-string component into the middle of the translation file.Your translation file should look something like thisNow we need to map the root element of the source 'singleString' to the XMLTYPE of the database adapter, applying the function get-content-as-string.To do this drag the element singleString to the left side of the function get-content-as-string and drag the right side of the get-content-as-string to the XMLFILE element of the database adapter so the mapping looks like this. You have now completed the SOA Component you can now save your work, deploy and test.When you deploy I have assumed that you have the correct database configurations in the WebLogic Console based on the connection you setup connecting to the Stored Procedure. 7. Testing the ApplicationOpen Enterprise Manager and navigate to the TestXMLTYPE Composite and click the Test button. Load some dummy variables in the Input Arguments and click the 'Test Web Service' buttonOnce completed you can run a SQL statement to check the install. In this instance I have just used JDeveloper and opened a SQL WorksheetSQL Statement Normal 0 false false false MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} select * from xml_example_tbl; Result, you should see the full payload in the result.

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  • FAQ: GridView Calculation with JavaScript

    - by Vincent Maverick Durano
    In my previous post I wrote a simple demo on how to Calculate Totals in GridView and Display it in the Footer. Basically what it does is it calculates the total amount by typing into the TextBox and display the grand total in the footer of the GridView and basically it was a server side implemenation.  Many users in the forums are asking how to do the same thing without postbacks and how to calculate both amount and total amount together. In this post I will demonstrate how to do this using JavaScript. To get started let's go ahead and set up the form. Just for the simplicity of this demo I just set up the form like this:   <asp:gridview ID="GridView1" runat="server" ShowFooter="true" AutoGenerateColumns="false"> <Columns> <asp:BoundField DataField="RowNumber" HeaderText="Row Number" /> <asp:BoundField DataField="Description" HeaderText="Item Description" /> <asp:TemplateField HeaderText="Item Price"> <ItemTemplate> <asp:Label ID="LBLPrice" runat="server" Text='<%# Eval("Price") %>'></asp:Label> </ItemTemplate> </asp:TemplateField> <asp:TemplateField HeaderText="Quantity"> <ItemTemplate> <asp:TextBox ID="TXTQty" runat="server"></asp:TextBox> </ItemTemplate> <FooterTemplate> <b>Total Amount:</b> </FooterTemplate> </asp:TemplateField> <asp:TemplateField HeaderText="Sub-Total"> <ItemTemplate> <asp:Label ID="LBLSubTotal" runat="server"></asp:Label> </ItemTemplate> <FooterTemplate> <asp:Label ID="LBLTotal" runat="server" ForeColor="Green"></asp:Label> </FooterTemplate> </asp:TemplateField> </Columns> </asp:gridview>   As you can see there's no fancy about the mark up above. It just a standard GridView with BoundFields and TemplateFields on it. Now just for the purpose of this demo I just use a dummy data for populating the GridView. Here's the code below:   public partial class GridCalculation : System.Web.UI.Page { private void BindDummyDataToGrid() { DataTable dt = new DataTable(); DataRow dr = null; dt.Columns.Add(new DataColumn("RowNumber", typeof(string))); dt.Columns.Add(new DataColumn("Description", typeof(string))); dt.Columns.Add(new DataColumn("Price", typeof(string))); dr = dt.NewRow(); dr["RowNumber"] = 1; dr["Description"] = "Nike"; dr["Price"] = "1000"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 2; dr["Description"] = "Converse"; dr["Price"] = "800"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 3; dr["Description"] = "Adidas"; dr["Price"] = "500"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 4; dr["Description"] = "Reebok"; dr["Price"] = "750"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 5; dr["Description"] = "Vans"; dr["Price"] = "1100"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 6; dr["Description"] = "Fila"; dr["Price"] = "200"; dt.Rows.Add(dr); //Bind the Gridview GridView1.DataSource = dt; GridView1.DataBind(); } protected void Page_Load(object sender, EventArgs e) { if (!IsPostBack) { BindDummyDataToGrid(); } } }   Now try to run the page. The output should look something like below: The Client-Side Calculation Here's the code for the GridView calculation:   <script type="text/javascript"> function CalculateTotals() { var gv = document.getElementById("<%= GridView1.ClientID %>"); var tb = gv.getElementsByTagName("input"); var lb = gv.getElementsByTagName("span"); var sub = 0; var total = 0; var indexQ = 1; var indexP = 0; for (var i = 0; i < tb.length; i++) { if (tb[i].type == "text") { sub = parseFloat(lb[indexP].innerHTML) * parseFloat(tb[i].value); if (isNaN(sub)) { lb[i + indexQ].innerHTML = ""; sub = 0; } else { lb[i + indexQ].innerHTML = sub; } indexQ++; indexP = indexP + 2; total += parseFloat(sub); } } lb[lb.length -1].innerHTML = total; } </script>   The code above calculates the sub-total by multiplying the price and the quantity and at the same time calculates the total amount  by adding the sub-total values. Now you can simply call the JavaScript function above like this:   <ItemTemplate> <asp:TextBox ID="TXTQty" runat="server" onkeyup="CalculateTotals();"></asp:TextBox> </ItemTemplate>   Running the code above will display something like below: That's it! I hope someone find this post useful! Technorati Tags: ASP.NET,JavaScript,GridView,TipsTricks

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  • Launching a WPF Window in a Separate Thread, Part 1

    - by Reed
    Typically, I strongly recommend keeping the user interface within an application’s main thread, and using multiple threads to move the actual “work” into background threads.  However, there are rare times when creating a separate, dedicated thread for a Window can be beneficial.  This is even acknowledged in the MSDN samples, such as the Multiple Windows, Multiple Threads sample.  However, doing this correctly is difficult.  Even the referenced MSDN sample has major flaws, and will fail horribly in certain scenarios.  To ease this, I wrote a small class that alleviates some of the difficulties involved. The MSDN Multiple Windows, Multiple Threads Sample shows how to launch a new thread with a WPF Window, and will work in most cases.  The sample code (commented and slightly modified) works out to the following: // Create a thread Thread newWindowThread = new Thread(new ThreadStart( () => { // Create and show the Window Window1 tempWindow = new Window1(); tempWindow.Show(); // Start the Dispatcher Processing System.Windows.Threading.Dispatcher.Run(); })); // Set the apartment state newWindowThread.SetApartmentState(ApartmentState.STA); // Make the thread a background thread newWindowThread.IsBackground = true; // Start the thread newWindowThread.Start(); .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; } This sample creates a thread, marks it as single threaded apartment state, and starts the Dispatcher on that thread. That is the minimum requirements to get a Window displaying and handling messages correctly, but, unfortunately, has some serious flaws. The first issue – the created thread will run continuously until the application shuts down, given the code in the sample.  The problem is that the ThreadStart delegate used ends with running the Dispatcher.  However, nothing ever stops the Dispatcher processing.  The thread was created as a Background thread, which prevents it from keeping the application alive, but the Dispatcher will continue to pump dispatcher frames until the application shuts down. In order to fix this, we need to call Dispatcher.InvokeShutdown after the Window is closed.  This would require modifying the above sample to subscribe to the Window’s Closed event, and, at that point, shutdown the Dispatcher: // Create a thread Thread newWindowThread = new Thread(new ThreadStart( () => { Window1 tempWindow = new Window1(); // When the window closes, shut down the dispatcher tempWindow.Closed += (s,e) => Dispatcher.CurrentDispatcher.BeginInvokeShutdown(DispatcherPriority.Background); tempWindow.Show(); // Start the Dispatcher Processing System.Windows.Threading.Dispatcher.Run(); })); // Setup and start thread as before This eliminates the first issue.  Now, when the Window is closed, the new thread’s Dispatcher will shut itself down, which in turn will cause the thread to complete. The above code will work correctly for most situations.  However, there is still a potential problem which could arise depending on the content of the Window1 class.  This is particularly nasty, as the code could easily work for most windows, but fail on others. The problem is, at the point where the Window is constructed, there is no active SynchronizationContext.  This is unlikely to be a problem in most cases, but is an absolute requirement if there is code within the constructor of Window1 which relies on a context being in place. While this sounds like an edge case, it’s fairly common.  For example, if a BackgroundWorker is started within the constructor, or a TaskScheduler is built using TaskScheduler.FromCurrentSynchronizationContext() with the expectation of synchronizing work to the UI thread, an exception will be raised at some point.  Both of these classes rely on the existence of a proper context being installed to SynchronizationContext.Current, which happens automatically, but not until Dispatcher.Run is called.  In the above case, SynchronizationContext.Current will return null during the Window’s construction, which can cause exceptions to occur or unexpected behavior. Luckily, this is fairly easy to correct.  We need to do three things, in order, prior to creating our Window: Create and initialize the Dispatcher for the new thread manually Create a synchronization context for the thread which uses the Dispatcher Install the synchronization context Creating the Dispatcher is quite simple – The Dispatcher.CurrentDispatcher property gets the current thread’s Dispatcher and “creates a new Dispatcher if one is not already associated with the thread.”  Once we have the correct Dispatcher, we can create a SynchronizationContext which uses the dispatcher by creating a DispatcherSynchronizationContext.  Finally, this synchronization context can be installed as the current thread’s context via SynchronizationContext.SetSynchronizationContext.  These three steps can easily be added to the above via a single line of code: // Create a thread Thread newWindowThread = new Thread(new ThreadStart( () => { // Create our context, and install it: SynchronizationContext.SetSynchronizationContext( new DispatcherSynchronizationContext( Dispatcher.CurrentDispatcher)); Window1 tempWindow = new Window1(); // When the window closes, shut down the dispatcher tempWindow.Closed += (s,e) => Dispatcher.CurrentDispatcher.BeginInvokeShutdown(DispatcherPriority.Background); tempWindow.Show(); // Start the Dispatcher Processing System.Windows.Threading.Dispatcher.Run(); })); // Setup and start thread as before This now forces the synchronization context to be in place before the Window is created and correctly shuts down the Dispatcher when the window closes. However, there are quite a few steps.  In my next post, I’ll show how to make this operation more reusable by creating a class with a far simpler API…

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  • Design for Vacation Tracking System

    - by Aaronaught
    I have been tasked with developing a system for tracking our company's paid time-off (vacation, sick days, etc.) At the moment we are using an Excel spreadsheet on a shared network drive, and it works pretty well, but we are concerned that we won't be able to "trust" employees forever and sometimes we run into locking issues when two people try to open the spreadsheet at once. So we are trying to build something a little more robust. I would like some input on this design in terms of maintainability, scalability, extensibility, etc. It's a pretty simple workflow we need to represent right now: I started with a basic MS Access schema like this: Employees (EmpID int, EmpName varchar(50), AllowedDays int) Vacations (VacationID int, EmpID int, BeginDate datetime, EndDate datetime) But we don't want to spend a lot of time building a schema and database like this and have to change it later, so I think I am going to go with something that will be easier to expand through configuration. Right now the vacation table has this schema: Vacations (VacationID int, PropName varchar(50), PropValue varchar(50)) And the table will be populated with data like this: VacationID | PropName | PropValue -----------+--------------+------------------ 1 | EmpID | 4 1 | EmpName | James Jones 1 | Reason | Vacation 1 | BeginDate | 2/24/2010 1 | EndDate | 2/30/2010 1 | Destination | Spectate Swamp 2 | ... | ... I think this is a pretty good, extensible design, we can easily add new properties to the vacation like the destination or maybe approval status, etc. I wasn't too sure how to go about managing the database of valid properties, I thought of putting them in a separate PropNames table but it gets complicated to manage all the different data types and people say that you shouldn't put CLR type names into a SQL database, so I decided to use XML instead, here is the schema: <VacationProperties> <PropertyNames>EmpID,EmpName,Reason,BeginDate,EndDate,Destination</PropertyNames> <PropertyTypes>System.Int32,System.String,System.String,System.DateTime,System.DateTime,System.String</PropertyTypes> <PropertiesRequired>true,true,false,true,true,false</PropertiesRequired> </VacationProperties> I might need more fields than that, I'm not completely sure. I'm parsing the XML like this (would like some feedback on the parsing code): string xml = File.ReadAllText("properties.xml"); Match m = Regex.Match(xml, "<(PropertyNames)>(.*?)</PropertyNames>"; string[] pn = m.Value.Split(','); // do the same for PropertyTypes, PropertiesRequired Then I use the following code to persist configuration changes to the database: string sql = "DROP TABLE VacationProperties"; sql = sql + " CREATE TABLE VacationProperties "; sql = sql + "(PropertyName varchar(100), PropertyType varchar(100) "; sql = sql + "IsRequired varchar(100))"; for (int i = 0; i < pn.Length; i++) { sql = sql + " INSERT VacationProperties VALUES (" + pn[i] + "," + pt[i] + "," + pv[i] + ")"; } // GlobalConnection is a singleton new SqlCommand(sql, GlobalConnection.Instance).ExecuteReader(); So far so good, but after a few days of this I then realized that a lot of this was just a more specific kind of a generic workflow which could be further abstracted, and instead of writing all of this boilerplate plumbing code I could just come up with a workflow and plug it into a workflow engine like Windows Workflow Foundation and have the users configure it: In order to support routing these configurations throw the workflow system, it seemed natural to implement generic XML Web Services for this instead of just using an XML file as above. I've used this code to implement the Web Services: public class VacationConfigurationService : WebService { [WebMethod] public void UpdateConfiguration(string xml) { // Above code goes here } } Which was pretty easy, although I'm still working on a way to validate that XML against some kind of schema as there's no error-checking yet. I also created a few different services for other operations like VacationSubmissionService, VacationReportService, VacationDataService, VacationAuthenticationService, etc. The whole Service Oriented Architecture looks like this: And because the workflow itself might change, I have been working on a way to integrate the WF workflow system with MS Visio, which everybody at the office already knows how to use so they could make changes pretty easily. We have a diagram that looks like the following (it's kind of hard to read but the main items are Activities, Authenticators, Validators, Transformers, Processors, and Data Connections, they're all analogous to the services in the SOA diagram above). The requirements for this system are: (Note - I don't control these, they were given to me by management) Main workflow must interface with Excel spreadsheet, probably through VBA macros (to ease the transition to the new system) Alerts should integrate with MS Outlook, Lotus Notes, and SMS (text messages). We also want to interface it with the company Voice Mail system but that is not a "hard" requirement. Performance requirements: Must handle 250,000 Transactions Per Second Should be able to handle up to 20,000 employees (right now we have 3) 99.99% uptime ("four nines") expected Must be secure against outside hacking, but users cannot be required to enter a username/password. Platforms: Must support Windows XP/Vista/7, Linux, iPhone, Blackberry, DOS 2.0, VAX, IRIX, PDP-11, Apple IIc. Time to complete: 6 to 8 weeks. My questions are: Is this a good design for the system so far? Am I using all of the recommended best practices for these technologies? How do I integrate the Visio diagram above with the Windows Workflow Foundation to call the ConfigurationService and persist workflow changes? Am I missing any important components? Will this be extensible enough to support any scenario via end-user configuration? Will the system scale to the above performance requirements? Will we need any expensive hardware to run it? Are there any "gotchas" I should know about with respect to cross-platform compatibility? For example would it be difficult to convert this to an iPhone app? How long would you expect this to take? (We've dedicated 1 week for testing so I'm thinking maybe 5 weeks?)

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  • Highlight Row in GridView with Colored Columns

    - by Vincent Maverick Durano
    I wrote a blog post a while back before here that demonstrate how to highlight a GridView row on mouseover and as you can see its very easy to highlight rows in GridView. One of my colleague uses the same technique for implemeting gridview row highlighting but the problem is that if a Column has background color on it that cell will not be highlighted anymore. To make it more clear then let's build up a sample application. ASPX:   1: <asp:GridView runat="server" id="GridView1" onrowcreated="GridView1_RowCreated" 2: onrowdatabound="GridView1_RowDataBound"> 3: </asp:GridView>   CODE BEHIND:   1: private DataTable FillData() { 2:   3: DataTable dt = new DataTable(); 4: DataRow dr = null; 5:   6: //Create DataTable columns 7: dt.Columns.Add(new DataColumn("RowNumber", typeof(string))); 8: dt.Columns.Add(new DataColumn("Col1", typeof(string))); 9: dt.Columns.Add(new DataColumn("Col2", typeof(string))); 10: dt.Columns.Add(new DataColumn("Col3", typeof(string))); 11:   12: //Create Row for each columns 13: dr = dt.NewRow(); 14: dr["RowNumber"] = 1; 15: dr["Col1"] = "A"; 16: dr["Col2"] = "B"; 17: dr["Col3"] = "C"; 18: dt.Rows.Add(dr); 19:   20: dr = dt.NewRow(); 21: dr["RowNumber"] = 2; 22: dr["Col1"] = "AA"; 23: dr["Col2"] = "BB"; 24: dr["Col3"] = "CC"; 25: dt.Rows.Add(dr); 26:   27: dr = dt.NewRow(); 28: dr["RowNumber"] = 3; 29: dr["Col1"] = "A"; 30: dr["Col2"] = "B"; 31: dr["Col3"] = "CC"; 32: dt.Rows.Add(dr); 33:   34: dr = dt.NewRow(); 35: dr["RowNumber"] = 4; 36: dr["Col1"] = "A"; 37: dr["Col2"] = "B"; 38: dr["Col3"] = "CC"; 39: dt.Rows.Add(dr); 40:   41: dr = dt.NewRow(); 42: dr["RowNumber"] = 5; 43: dr["Col1"] = "A"; 44: dr["Col2"] = "B"; 45: dr["Col3"] = "CC"; 46: dt.Rows.Add(dr); 47:   48: return dt; 49: } 50:   51: protected void Page_Load(object sender, EventArgs e) { 52: if (!IsPostBack) { 53: GridView1.DataSource = FillData(); 54: GridView1.DataBind(); 55: } 56: }   As you can see there's nothing fancy in the code above. It just contain a method that fills a DataTable with a dummy data on it. Now here's the code for row highlighting:   1: protected void GridView1_RowCreated(object sender, GridViewRowEventArgs e) { 2: //Set Background Color for Columns 1 and 3 3: e.Row.Cells[1].BackColor = System.Drawing.Color.Beige; 4: e.Row.Cells[3].BackColor = System.Drawing.Color.Red; 5:   6: //Attach onmouseover and onmouseout for row highlighting 7: e.Row.Attributes.Add("onmouseover", "this.style.backgroundColor='Blue'"); 8: e.Row.Attributes.Add("onmouseout", "this.style.backgroundColor=''"); 9: }   Running the code above will show something like this in the browser: On initial load: On mouseover of GridView row:   Noticed that Col1 and Col3 are not highlighted. Why? the reason is that Col1 and Col3 cells has background color set on it and we only highlight the rows (TR) and not the columns (TD) that's why on mouseover only the rows will be highlighted. To fix the issue we will create a javascript method that would remove the background color of the columns when highlighting a row and on mouseout set back the original color that is set on Col1 and Col3. Here are the codes below: JavaScript   1: <script type="text/javascript"> 2: function HighLightRow(rowIndex, colIndex,colIndex2, flag) { 3: var gv = document.getElementById("<%= GridView1.ClientID %>"); 4: var selRow = gv.rows[rowIndex]; 5: if (rowIndex > 0) { 6: if (flag == "sel") { 7: gv.rows[rowIndex].style.backgroundColor = 'Blue'; 8: gv.rows[rowIndex].style.color = "White"; 9: gv.rows[rowIndex].cells[colIndex].style.backgroundColor = ''; 10: gv.rows[rowIndex].cells[colIndex2].style.backgroundColor = ''; 11: } 12: else { 13: gv.rows[rowIndex].style.backgroundColor = ''; 14: gv.rows[rowIndex].style.color = "Black"; 15: gv.rows[rowIndex].cells[colIndex].style.backgroundColor = 'Beige'; 16: gv.rows[rowIndex].cells[colIndex2].style.backgroundColor = 'Red'; 17: } 18: } 19: } 20: </script>   The HighLightRow method is a javascript function that accepts four (4) parameters which are the rowIndex,colIndex,colIndex2 and the flag. The rowIndex is the current row index of the selected row in GridView. The colIndex is the index of Col1 and colIndex2 is the index of col3. We are passing these index because these columns has background color on it and we need to toggle its backgroundcolor when highlighting the row in GridView. Finally the flag is something that would determine if its selected or not. Now here's the code for calling the JavaScript function above.     1: protected void GridView1_RowCreated(object sender, GridViewRowEventArgs e) { 2:   3: //Set Background Color for Columns 1 and 3 4: e.Row.Cells[1].BackColor = System.Drawing.Color.Beige; 5: e.Row.Cells[3].BackColor = System.Drawing.Color.Red; 6:   7: //Attach onmouseover and onmouseout for row highlighting 8: //and call the HighLightRow method with the required parameters 9: int index = e.Row.RowIndex + 1; 10: e.Row.Attributes.Add("onmouseover", "HighLightRow(" + index + "," + 1 + "," + 3 + ",'sel')"); 11: e.Row.Attributes.Add("onmouseout", "HighLightRow(" + index + "," + 1 + "," + 3 + ",'dsel')"); 12: 13: }   Running the code above will display something like this: On initial load:   On mouseover of GridView row:   That's it! I hope someone find this post useful!

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  • Using Radio Button in GridView with Validation

    - by Vincent Maverick Durano
    A developer is asking how to select one radio button at a time if the radio button is inside the GridView.  As you may know setting the group name attribute of radio button will not work if the radio button is located within a Data Representation control like GridView. This because the radio button inside the gridview bahaves differentely. Since a gridview is rendered as table element , at run time it will assign different "name" to each radio button. Hence you are able to select multiple rows. In this post I'm going to demonstrate how select one radio button at a time in gridview and add a simple validation on it. To get started let's go ahead and fire up visual studio and the create a new web application / website project. Add a WebForm and then add gridview. The mark up would look something like this: <asp:GridView ID="GridView1" runat="server" AutoGenerateColumns="false" > <Columns> <asp:TemplateField> <ItemTemplate> <asp:RadioButton ID="rb" runat="server" /> </ItemTemplate> </asp:TemplateField> <asp:BoundField DataField="RowNumber" HeaderText="Row Number" /> <asp:BoundField DataField="Col1" HeaderText="First Column" /> <asp:BoundField DataField="Col2" HeaderText="Second Column" /> </Columns> </asp:GridView> Noticed that I've added a templatefield column so that we can add the radio button there. Also I have set up some BoundField columns and set the DataFields as RowNumber, Col1 and Col2. These columns are just dummy columns and i used it for the simplicity of this example. Now where these columns came from? These columns are created by hand at the code behind file of the ASPX. Here's the code below: private DataTable FillData() { DataTable dt = new DataTable(); DataRow dr = null; //Create DataTable columns dt.Columns.Add(new DataColumn("RowNumber", typeof(string))); dt.Columns.Add(new DataColumn("Col1", typeof(string))); dt.Columns.Add(new DataColumn("Col2", typeof(string))); //Create Row for each columns dr = dt.NewRow(); dr["RowNumber"] = 1; dr["Col1"] = "A"; dr["Col2"] = "B"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 2; dr["Col1"] = "AA"; dr["Col2"] = "BB"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 3; dr["Col1"] = "A"; dr["Col2"] = "B"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 4; dr["Col1"] = "A"; dr["Col2"] = "B"; dt.Rows.Add(dr); dr = dt.NewRow(); dr["RowNumber"] = 5; dr["Col1"] = "A"; dr["Col2"] = "B"; dt.Rows.Add(dr); return dt; } And here's the code for binding the GridView with the dummy data above. protected void Page_Load(object sender, EventArgs e) { if (!IsPostBack) { GridView1.DataSource = FillData(); GridView1.DataBind(); } } Okay we have now a GridView data with a radio button on each row. Now lets go ahead and switch back to ASPX mark up. In this example I'm going to use a JavaScript for validating the radio button to select one radio button at a time. Here's the javascript code below: function CheckOtherIsCheckedByGVID(rb) { var isChecked = rb.checked; var row = rb.parentNode.parentNode; if (isChecked) { row.style.backgroundColor = '#B6C4DE'; row.style.color = 'black'; } var currentRdbID = rb.id; parent = document.getElementById("<%= GridView1.ClientID %>"); var items = parent.getElementsByTagName('input'); for (i = 0; i < items.length; i++) { if (items[i].id != currentRdbID && items[i].type == "radio") { if (items[i].checked) { items[i].checked = false; items[i].parentNode.parentNode.style.backgroundColor = 'white'; items[i].parentNode.parentNode.style.color = '#696969'; } } } } The function above sets the row of the current selected radio button's style to determine that the row is selected and then loops through the radio buttons in the gridview and then de-select the previous selected radio button and set the row style back to its default. You can then call the javascript function above at onlick event of radio button like below: <asp:RadioButton ID="rb" runat="server" onclick="javascript:CheckOtherIsCheckedByGVID(this);" /> Here's the output below: On Load: After Selecting a Radio Button: As you have noticed, on initial load there's no default selected radio in the GridView. Now let's add a simple validation for that. We will basically display an error message if a user clicks a button that triggers a postback without selecting  a radio button in the GridView. Here's the javascript for the validation: function ValidateRadioButton(sender, args) { var gv = document.getElementById("<%= GridView1.ClientID %>"); var items = gv.getElementsByTagName('input'); for (var i = 0; i < items.length ; i++) { if (items[i].type == "radio") { if (items[i].checked) { args.IsValid = true; return; } else { args.IsValid = false; } } } } The function above loops through the rows in gridview and find all the radio buttons within it. It will then check each radio button checked property. If a radio is checked then set IsValid to true else set it to false.  The reason why I'm using IsValid is because I'm using the ASP validator control for validation. Now add the following mark up below under the GridView declaration: <br /> <asp:Label ID="lblMessage" runat="server" /> <br /> <asp:Button ID="btn" runat="server" Text="POST" onclick="btn_Click" ValidationGroup="GroupA" /> <asp:CustomValidator ID="CustomValidator1" runat="server" ErrorMessage="Please select row in the grid." ClientValidationFunction="ValidateRadioButton" ValidationGroup="GroupA" style="display:none"></asp:CustomValidator> <asp:ValidationSummary ID="ValidationSummary1" runat="server" ValidationGroup="GroupA" HeaderText="Error List:" DisplayMode="BulletList" ForeColor="Red" /> And then at Button Click event add this simple code below just to test if  the validation works: protected void btn_Click(object sender, EventArgs e) { lblMessage.Text = "Postback at: " + DateTime.Now.ToString("hh:mm:ss tt"); } Here's the output below that you can see in the browser:   That's it! I hope someone find this post useful! Technorati Tags: ASP.NET,JavaScript,GridView

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  • Metro: Grouping Items in a ListView Control

    - by Stephen.Walther
    The purpose of this blog entry is to explain how you can group list items when displaying the items in a WinJS ListView control. In particular, you learn how to group a list of products by product category. Displaying a grouped list of items in a ListView control requires completing the following steps: Create a Grouped data source from a List data source Create a Grouped Header Template Declare the ListView control so it groups the list items Creating the Grouped Data Source Normally, you bind a ListView control to a WinJS.Binding.List object. If you want to render list items in groups, then you need to bind the ListView to a grouped data source instead. The following code – contained in a file named products.js — illustrates how you can create a standard WinJS.Binding.List object from a JavaScript array and then return a grouped data source from the WinJS.Binding.List object by calling its createGrouped() method: (function () { "use strict"; // Create List data source var products = new WinJS.Binding.List([ { name: "Milk", price: 2.44, category: "Beverages" }, { name: "Oranges", price: 1.99, category: "Fruit" }, { name: "Wine", price: 8.55, category: "Beverages" }, { name: "Apples", price: 2.44, category: "Fruit" }, { name: "Steak", price: 1.99, category: "Other" }, { name: "Eggs", price: 2.44, category: "Other" }, { name: "Mushrooms", price: 1.99, category: "Other" }, { name: "Yogurt", price: 2.44, category: "Other" }, { name: "Soup", price: 1.99, category: "Other" }, { name: "Cereal", price: 2.44, category: "Other" }, { name: "Pepsi", price: 1.99, category: "Beverages" } ]); // Create grouped data source var groupedProducts = products.createGrouped( function (dataItem) { return dataItem.category; }, function (dataItem) { return { title: dataItem.category }; }, function (group1, group2) { return group1.charCodeAt(0) - group2.charCodeAt(0); } ); // Expose the grouped data source WinJS.Namespace.define("ListViewDemos", { products: groupedProducts }); })(); Notice that the createGrouped() method requires three functions as arguments: groupKey – This function associates each list item with a group. The function accepts a data item and returns a key which represents a group. In the code above, we return the value of the category property for each product. groupData – This function returns the data item displayed by the group header template. For example, in the code above, the function returns a title for the group which is displayed in the group header template. groupSorter – This function determines the order in which the groups are displayed. The code above displays the groups in alphabetical order: Beverages, Fruit, Other. Creating the Group Header Template Whenever you create a ListView control, you need to create an item template which you use to control how each list item is rendered. When grouping items in a ListView control, you also need to create a group header template. The group header template is used to render the header for each group of list items. Here’s the markup for both the item template and the group header template: <div id="productTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> <div id="productGroupHeaderTemplate" data-win-control="WinJS.Binding.Template"> <div class="productGroupHeader"> <h1 data-win-bind="innerText: title"></h1> </div> </div> You should declare the two templates in the same file as you declare the ListView control – for example, the default.html file. Declaring the ListView Control The final step is to declare the ListView control. Here’s the required markup: <div data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource:ListViewDemos.products.dataSource, itemTemplate:select('#productTemplate'), groupDataSource:ListViewDemos.products.groups.dataSource, groupHeaderTemplate:select('#productGroupHeaderTemplate'), layout: {type: WinJS.UI.GridLayout} }"> </div> In the markup above, six properties of the ListView control are set when the control is declared. First the itemDataSource and itemTemplate are specified. Nothing new here. Next, the group data source and group header template are specified. Notice that the group data source is represented by the ListViewDemos.products.groups.dataSource property of the grouped data source. Finally, notice that the layout of the ListView is changed to Grid Layout. You are required to use Grid Layout (instead of the default List Layout) when displaying grouped items in a ListView. Here’s the entire contents of the default.html page: <!DOCTYPE html> <html> <head> <meta charset="utf-8"> <title>ListViewDemos</title> <!-- WinJS references --> <link href="//Microsoft.WinJS.0.6/css/ui-dark.css" rel="stylesheet"> <script src="//Microsoft.WinJS.0.6/js/base.js"></script> <script src="//Microsoft.WinJS.0.6/js/ui.js"></script> <!-- ListViewDemos references --> <link href="/css/default.css" rel="stylesheet"> <script src="/js/default.js"></script> <script src="/js/products.js" type="text/javascript"></script> <style type="text/css"> .product { width: 200px; height: 100px; border: white solid 1px; font-size: x-large; } </style> </head> <body> <div id="productTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> <div id="productGroupHeaderTemplate" data-win-control="WinJS.Binding.Template"> <div class="productGroupHeader"> <h1 data-win-bind="innerText: title"></h1> </div> </div> <div data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource:ListViewDemos.products.dataSource, itemTemplate:select('#productTemplate'), groupDataSource:ListViewDemos.products.groups.dataSource, groupHeaderTemplate:select('#productGroupHeaderTemplate'), layout: {type: WinJS.UI.GridLayout} }"> </div> </body> </html> Notice that the default.html page includes a reference to the products.js file: <script src=”/js/products.js” type=”text/javascript”></script> The default.html page also contains the declarations of the item template, group header template, and ListView control. Summary The goal of this blog entry was to explain how you can group items in a ListView control. You learned how to create a grouped data source, a group header template, and declare a ListView so that it groups its list items.

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  • Establishing WebLogic Server HTTPS Trust of IIS Using a Microsoft Local Certificate Authority

    - by user647124
    Everyone agrees that self-signed and demo certificates for SSL and HTTPS should never be used in production and preferred not to be used elsewhere. Most self-signed and demo certificates are provided by vendors with the intention that they are used only to integrate within the same environment. In a vendor’s perfect world all application servers in a given enterprise are from the same vendor, which makes this lack of interoperability in a non-production environment an advantage. For us working in the real world, where not only do we not use a single vendor everywhere but have to make do with self-signed certificates for all but production, testing HTTPS between an IIS ASP.NET service provider and a WebLogic J2EE consumer application can be very frustrating to set up. It was for me, especially having found many blogs and discussion threads where various solutions were described but did not quite work and were all mostly similar but just a little bit different. To save both you and my future (who always seems to forget the hardest-won lessons) all of the pain and suffering, I am recording the steps that finally worked here for reference and sanity. How You Know You Need This The first cold clutches of dread that tells you it is going to be a long day is when you attempt to a WSDL published by IIS in WebLogic over HTTPS and you see the following: <Jul 30, 2012 2:51:31 PM EDT> <Warning> <Security> <BEA-090477> <Certificate chain received from myserver.mydomain.com - 10.555.55.123 was not trusted causing SSL handshake failure.> weblogic.wsee.wsdl.WsdlException: Failed to read wsdl file from url due to -- javax.net.ssl.SSLKeyException: [Security:090477]Certificate chain received from myserver02.mydomain.com - 10.555.55.123 was not trusted causing SSL handshake failure. The above is what started a three day sojourn into searching for a solution. Even people who had solved it before would tell me how they did, and then shrug when I demonstrated that the steps did not end in the success they claimed I would experience. Rather than torture you with the details of everything I did that did not work, here is what finally did work. Export the Certificates from IE First, take the offending WSDL URL and paste it into IE (if you have an internal Microsoft CA, you have IE, even if you don’t use it in favor of some other browser). To state the semi-obvious, if you received the error above there is a certificate configured for the IIS host of the service and the SSL port has been configured properly. Otherwise there would be a different error, usually about the site not found or connection failed. Once the WSDL loads, to the right of the address bar there will be a lock icon. Click the lock and then click View Certificates in the resulting dialog (if you do not have a lock icon but do have a Certificate Error message, see http://support.microsoft.com/kb/931850 for steps to install the certificate then you can continue from the point of finding the lock icon). Figure 1: View Certificates in IE Next, select the Details tab in the resulting dialog Figure 2: Use Certificate Details to Export Certificate Click Copy to File, then Next, then select the Base-64 encoded option for the format Figure 3: Select the Base-64 encoded option for the format For the sake of simplicity, I choose to save this to the root of the WebLogic domain. It will work from anywhere, but later you will need to type in the full path rather than just the certificate name if you save it elsewhere. Figure 4: Browse to Save Location Figure 5: Save the Certificate to the Domain Root for Convenience This is the point where I ran into some confusion. Some articles mentioned exporting the entire chain of certificates. This supposedly works for some types of certificates, or if you have a few other tools and the time to learn them. For the SSL experts out there, they already have these tools, know how to use them well, and should not be wasting their time reading this article meant for folks who just want to get things wired up and back to unit testing and development. For the rest of us, the easiest way to make sure things will work is to just export all the links in the chain individually and let WebLogic Server worry about re-assembling them into a chain (which it does quite nicely). While perhaps not the most elegant solution, the multi-step process is easy to repeat and uses only tools that are immediately available and require no learning curve. So… Next, go to Tools then Internet Options then the Content tab and click Certificates. Go to the Trust Root Certificate Authorities tab and find the certificate root for your Microsoft CA cert (look for the Issuer of the certificate you exported earlier). Figure 6: Trusted Root Certification Authorities Tab Export this one the same way as before, with a different name Figure 7: Use a Unique Name for Each Certificate Repeat this once more for the Intermediate Certificate tab. Import the Certificates to the WebLogic Domain Now, open an command prompt, navigate to [WEBLOGIC_DOMAIN_ROOT]\bin and execute setDomainEnv. You should then be in the root of the domain. If not, CD to the domain root. Assuming you saved the certificate in the domain root, execute the following: keytool -importcert -alias [ALIAS-1] -trustcacerts -file [FULL PATH TO .CER 1] -keystore truststore.jks -storepass [PASSWORD] An example with the variables filled in is: keytool -importcert -alias IIS-1 -trustcacerts -file microsftcert.cer -keystore truststore.jks -storepass password After several lines out output you will be prompted with: Trust this certificate? [no]: The correct answer is ‘yes’ (minus the quotes, of course). You’ll you know you were successful if the response is: Certificate was added to keystore If not, check your typing, as that is generally the source of an error at this point. Repeat this for all three of the certificates you exported, changing the [ALIAS-1] and [FULL PATH TO .CER 1] value each time. For example: keytool -importcert -alias IIS-1 -trustcacerts -file microsftcert.cer -keystore truststore.jks -storepass password keytool -importcert -alias IIS-2 -trustcacerts -file microsftcertRoot.cer -keystore truststore.jks -storepass password keytool -importcert -alias IIS-3 -trustcacerts -file microsftcertIntermediate.cer -keystore truststore.jks -storepass password In the above we created a new JKS key store. You can re-use an existing one by changing the name of the JKS file to one you already have and change the password to the one that matches that JKS file. For the DemoTrust.jks  that is included with WebLogic the password is DemoTrustKeyStorePassPhrase. An example here would be: keytool -importcert -alias IIS-1 -trustcacerts -file microsoft.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase keytool -importcert -alias IIS-2 -trustcacerts -file microsoftRoot.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase keytool -importcert -alias IIS-2 -trustcacerts -file microsoftInter.cer -keystore DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase Whichever keystore you use, you can check your work with: keytool -list -keystore truststore.jks -storepass password Where “truststore.jks” and “password” can be replaced appropriately if necessary. The output will look something like this: Figure 8: Output from keytool -list -keystore Update the WebLogic Keystore Configuration If you used an existing keystore rather than creating a new one, you can restart your WebLogic Server and skip the rest of this section. For those of us who created a new one because that is the instructions we found online… Next, we need to tell WebLogic to use the JKS file (truststore.jks) we just created. Log in to the WebLogic Server Administration Console and navigate to Servers > AdminServer > Configuration > Keystores. Scroll down to “Custom Trust Keystore:” and change the value to “truststore.jks” and the value of “Custom Trust Keystore Passphrase:” and “Confirm Custom Trust Keystore Passphrase:” to the password you used when earlier, then save your changes. You will get a nice message similar to the following: Figure 9: To Be Safe, Restart Anyways The “No restarts are necessary” is somewhat of an exaggeration. If you want to be able to use the keystore you may need restart the server(s). To save myself aggravation, I always do. Your mileage may vary. Conclusion That should get you there. If there are some erroneous steps included for your situation in particular, I will offer up a semi-apology as the process described above does not take long at all and if there is one step that could be dropped from it, is still much faster than trying to figure this out from other sources.

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