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  • Covariance and Contravariance in C#

    - by edalorzo
    I will start by saying that I am Java developer learning to program in C#. As such I do comparisons of what I know with what I am learning. I have been playing with C# generics for a few hours now, and I have been able to reproduce the same things I know in Java in C#, with the exception of a couple of examples using covariance and contravariance. The book I am reading is not very good in the subject. I will certainly seek more info on the web, but while I do that, perhaps you can help me find a C# implementation for the following Java code. An example is worth a thousand words, and I was hoping that by looking a good code sample I will be able to assimilate this more rapidly. Covariance In Java I can do something like this: public static double sum(List<? extends Number> numbers) { double summation = 0.0; for(Number number : numbers){ summation += number.doubleValue(); } return summation; } I can use this code as follows: List<Integer> myInts = asList(1,2,3,4,5); List<Double> myDoubles = asList(3.14, 5.5, 78.9); List<Long> myLongs = asList(1L, 2L, 3L); double result = 0.0; result = sum(myInts); result = sum(myDoubles) result = sum(myLongs); Now I did discover that C# supports covariance/contravariance only on interfaces and as long as they have been explicitly declared to do so (out). I think I was not able to reproduce this case, because I could not find a common ancestor of all numbers, but I believe that I could have used IEnumerable to implement such thing if a common ancestor exists. Since IEnumerable is a covariant type. Right? Any thoughts on how to implement the list above? Just point me into the right direction. Is there any common ancestor of all numeric types? Contravariance The contravariance example I tried was the following. In Java I can do this to copy one list into another. public static void copy(List<? extends Number> source, List<? super Number> destiny){ for(Number number : source) { destiny.add(number); } } Then I could use it with contravariant types as follows: List<Object> anything = new ArrayList<Object>(); List<Integer> myInts = asList(1,2,3,4,5); copy(myInts, anything); My basic problem, trying to implement this in C# is that I could not find an interface that was both covariant and contravariant at the same time, as it is case of List in my example above. Maybe it can be done with two different interface in C#. Any thoughts on how to implement this? Thank you very much to everyone for any answers you can contribute. I am pretty sure I will learn a lot from any example you can provide.

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  • Game AI: Pattern for implementing Sense-Think-Act components?

    - by Rosarch
    I'm developing a game. Each entity in the game is a GameObject. Each GameObject is composed of a GameObjectController, GameObjectModel, and GameObjectView. (Or inheritants thereof.) For NPCs, the GameObjectController is split into: IThinkNPC: reads current state and makes a decision about what to do IActNPC: updates state based on what needs to be done ISenseNPC: reads current state to answer world queries (eg "am I being in the shadows?") My question: Is this ok for the ISenseNPC interface? public interface ISenseNPC { // ... /// <summary> /// True if `dest` is a safe point to which to retreat. /// </summary> /// <param name="dest"></param> /// <param name="angleToThreat"></param> /// <param name="range"></param> /// <returns></returns> bool IsSafeToRetreat(Vector2 dest, float angleToThreat, float range); /// <summary> /// Finds a new location to which to retreat. /// </summary> /// <param name="angleToThreat"></param> /// <returns></returns> Vector2 newRetreatDest(float angleToThreat); /// <summary> /// Returns the closest LightSource that illuminates the NPC. /// Null if the NPC is not illuminated. /// </summary> /// <returns></returns> ILightSource ClosestIlluminatingLight(); /// <summary> /// True if the NPC is sufficiently far away from target. /// Assumes that target is the only entity it could ever run from. /// </summary> /// <returns></returns> bool IsSafeFromTarget(); } None of the methods take any parameters. Instead, the implementation is expected to maintain a reference to the relevant GameObjectController and read that. However, I'm now trying to write unit tests for this. Obviously, it's necessary to use mocking, since I can't pass arguments directly. The way I'm doing it feels really brittle - what if another implementation comes along that uses the world query utilities in a different way? Really, I'm not testing the interface, I'm testing the implementation. Poor. The reason I used this pattern in the first place was to keep IThinkNPC implementation code clean: public BehaviorState RetreatTransition(BehaviorState currentBehavior) { if (sense.IsCollidingWithTarget()) { NPCUtils.TraceTransitionIfNeeded(ToString(), BehaviorState.ATTACK.ToString(), "is colliding with target"); return BehaviorState.ATTACK; } if (sense.IsSafeFromTarget() && sense.ClosestIlluminatingLight() == null) { return BehaviorState.WANDER; } if (sense.ClosestIlluminatingLight() != null && sense.SeesTarget()) { NPCUtils.TraceTransitionIfNeeded(ToString(), BehaviorState.ATTACK.ToString(), "collides with target"); return BehaviorState.CHASE; } return currentBehavior; } Perhaps the cleanliness isn't worth it, however. So, if ISenseNPC takes all the params it needs every time, I could make it static. Is there any problem with that?

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  • Languages and VMs: Features that are hard to optimize and why

    - by mrjoltcola
    I'm doing a survey of features in preparation for a research project. Name a mainstream language or language feature that is hard to optimize, and why the feature is or isn't worth the price paid, or instead, just debunk my theories below with anecdotal evidence. Before anyone flags this as subjective, I am asking for specific examples of languages or features, and ideas for optimization of these features, or important features that I haven't considered. Also, any references to implementations that prove my theories right or wrong. Top on my list of hard to optimize features and my theories (some of my theories are untested and are based on thought experiments): 1) Runtime method overloading (aka multi-method dispatch or signature based dispatch). Is it hard to optimize when combined with features that allow runtime recompilation or method addition. Or is it just hard, anyway? Call site caching is a common optimization for many runtime systems, but multi-methods add additional complexity as well as making it less practical to inline methods. 2) Type morphing / variants (aka value based typing as opposed to variable based) Traditional optimizations simply cannot be applied when you don't know if the type of someting can change in a basic block. Combined with multi-methods, inlining must be done carefully if at all, and probably only for a given threshold of size of the callee. ie. it is easy to consider inlining simple property fetches (getters / setters) but inlining complex methods may result in code bloat. The other issue is I cannot just assign a variant to a register and JIT it to the native instructions because I have to carry around the type info, or every variable needs 2 registers instead of 1. On IA-32 this is inconvenient, even if improved with x64's extra registers. This is probably my favorite feature of dynamic languages, as it simplifies so many things from the programmer's perspective. 3) First class continuations - There are multiple ways to implement them, and I have done so in both of the most common approaches, one being stack copying and the other as implementing the runtime to use continuation passing style, cactus stacks, copy-on-write stack frames, and garbage collection. First class continuations have resource management issues, ie. we must save everything, in case the continuation is resumed, and I'm not aware if any languages support leaving a continuation with "intent" (ie. "I am not coming back here, so you may discard this copy of the world"). Having programmed in the threading model and the contination model, I know both can accomplish the same thing, but continuations' elegance imposes considerable complexity on the runtime and also may affect cache efficienty (locality of stack changes more with use of continuations and co-routines). The other issue is they just don't map to hardware. Optimizing continuations is optimizing for the less-common case, and as we know, the common case should be fast, and the less-common cases should be correct. 4) Pointer arithmetic and ability to mask pointers (storing in integers, etc.) Had to throw this in, but I could actually live without this quite easily. My feelings are that many of the high-level features, particularly in dynamic languages just don't map to hardware. Microprocessor implementations have billions of dollars of research behind the optimizations on the chip, yet the choice of language feature(s) may marginalize many of these features (features like caching, aliasing top of stack to register, instruction parallelism, return address buffers, loop buffers and branch prediction). Macro-applications of micro-features don't necessarily pan out like some developers like to think, and implementing many languages in a VM ends up mapping native ops into function calls (ie. the more dynamic a language is the more we must lookup/cache at runtime, nothing can be assumed, so our instruction mix is made up of a higher percentage of non-local branching than traditional, statically compiled code) and the only thing we can really JIT well is expression evaluation of non-dynamic types and operations on constant or immediate types. It is my gut feeling that bytecode virtual machines and JIT cores are perhaps not always justified for certain languages because of this. I welcome your answers.

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  • TDD - beginner problems and stumbling blocks

    - by Noufal Ibrahim
    While I've written unit tests for most of the code I've done, I only recently got my hands on a copy of TDD by example by Kent Beck. I have always regretted certain design decisions I made since they prevented the application from being 'testable'. I read through the book and while some of it looks alien, I felt that I could manage it and decided to try it out on my current project which is basically a client/server system where the two pieces communicate via. USB. One on the gadget and the other on the host. The application is in Python. I started off and very soon got entangled in a mess of rewrites and tiny tests which I later figured didn't really test anything. I threw away most of them and and now have a working application for which the tests have all coagulated into just 2. Based on my experiences, I have a few questions which I'd like to ask. I gained some information from http://stackoverflow.com/questions/1146218/new-to-tdd-are-there-sample-applications-with-tests-to-show-how-to-do-tdd but have some specific questions which I'd like answers to/discussion on. Kent Beck uses a list which he adds to and strikes out from to guide the development process. How do you make such a list? I initially had a few items like "server should start up", "server should abort if channel is not available" etc. but they got mixed and finally now, it's just something like "client should be able to connect to server" (which subsumed server startup etc.). How do you handle rewrites? I initially selected a half duplex system based on named pipes so that I could develop the application logic on my own machine and then later add the USB communication part. It them moved to become a socket based thing and then moved from using raw sockets to using the Python SocketServer module. Each time things changed, I found that I had to rewrite considerable parts of the tests which was annoying. I'd figured that the tests would be a somewhat invariable guide during my development. They just felt like more code to handle. I needed a client and a server to communicate through the channel to test either side. I could mock one of the sides to test the other but then the whole channel wouldn't be tested and I worry that I'd miss that. This detracted from the whole red/green/refactor rhythm. Is this just lack of experience or am I doing something wrong? The "Fake it till you make it" left me with a lot of messy code that I later spent a lot of time to refactor and clean up. Is this the way things work? At the end of the session, I now have my client and server running with around 3 or 4 unit tests. It took me around a week to do it. I think I could have done it in a day if I were using the unit tests after code way. I fail to see the gain. I'm looking for comments and advice from people who have implemented large non trivial projects completely (or almost completely) using this methodology. It makes sense to me to follow the way after I have something already running and want to add a new feature but doing it from scratch seems to tiresome and not worth the effort. P.S. : Please let me know if this should be community wiki and I'll mark it like that. Update 0 : All the answers were equally helpful. I picked the one I did because it resonated with my experiences the most. Update 1: Practice Practice Practice!

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  • Help in building an 16 bit os

    - by Barshan Das
    I am trying to build an old 16 bit dos like os. My bootloader code: ; This is not my code. May be of Fritzos. I forgot the source. ORG 7c00h jmp Start drive db 0 msg db " Loader Initialization",0 msg2 db "ACos Loaded",0 print: lodsb cmp al, 0 je end mov ah, 0Eh int 10h jmp print end: ret Start: mov [ drive ], dl ; Get the floppy OS booted from ; Update the segment registers xor ax, ax ; XOR ax mov ds, ax ; Mov AX into DS mov si,msg call print ; Load Kernel. ResetFloppy: mov ax, 0x00 ; Select Floppy Reset BIOS Function mov dl, [ drive ] ; Select the floppy ADos booted from int 13h ; Reset the floppy drive jc ResetFloppy ; If there was a error, try again. ReadFloppy: mov bx, 0x9000 ; Load kernel at 9000h. mov ah, 0x02 ; Load disk data to ES:BX mov al, 17 ; Load two floppy head full's worth of data. mov ch, 0 ; First Cylinder mov cl, 2 ; Start at the 2nd Sector to load the Kernel mov dh, 0 ; Use first floppy head mov dl, [ drive ] ; Load from the drive kernel booted from. int 13h ; Read the floppy disk. jc ReadFloppy ; Error, try again. ; Clear text mode screen mov ax, 3 int 10h ;print starting message mov si,msg2 call print mov ax, 0x0 mov ss, ax mov sp, 0xFFFF jmp 9000h ; This part makes sure the bootsector is 512 bytes. times 510-($-$$) db 0 ;bootable sector signature dw 0xAA55 My example kernel code: asm(".code16\n"); void putchar(char); int main() { putchar('A'); return 0; } void putchar(char val) { asm("movb %0, %%al\n" "movb $0x0E, %%ah\n" "int $0x10\n" : :"r"(val) ) ; } This is how I compile it : nasm -f bin -o ./bin/boot.bin ./source/boot.asm gcc -nostdinc -fno-builtin -I./include -c -o ./bin/kernel.o ./source/kernel.c ld -Ttext=0x9000 -o ./bin/kernel.bin ./bin/kernel.o -e 0x0 dd if=/dev/zero of=./bin/empty.bin bs=1440K count=1 cat ./bin/boot.bin ./bin/kernel.bin ./bin/empty.bin|head -c 1440K > ./bin/os rm ./bin/empty.bin and I run it in virtual machine. When I make the putchar function ( in kernel code ) for constant value ....i.e like this: void putchar() { char val = 'A'; asm("movb %0, %%al\n" "movb $0x0E, %%ah\n" "int $0x10\n" : :"r"(val) ) ; } then it works fine. But when I pass argument to it ( That is in the previous code ) , then it prints a space for any character. What should I do?

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  • hello-1.mod.c:14: warning: missing initializer (near initialization for '__this_module.arch.unw_sec_init')

    - by Sompom
    I am trying to write a module for an sbc1651. Since the device is ARM, this requires a cross-compile. As a start, I am trying to compile the "Hello Kernel" module found here. This compiles fine on my x86 development system, but when I try to cross-compile I get the below error. /home/developer/HelloKernel/hello-1.mod.c:14: warning: missing initializer /home/developer/HelloKernel/hello-1.mod.c:14: warning: (near initialization for '__this_module.arch.unw_sec_init') Since this is in the .mod.c file, which is autogenerated I have no idea what's going on. The mod.c file seems to be generated by the module.h file. As far as I can tell, the relevant parts are the same between my x86 system's module.h and the arm kernel header's module.h. Adding to my confusion, this problem is either not googleable (by me...) or hasn't happened to anyone before. Or I'm just doing something clueless that anyone with any sense wouldn't do. The cross-compiler I'm using was supplied by Freescale (I think). I suppose it could be a problem with the compiler. Would it be worth trying to build the toolchain myself? Obviously, since this is a warning, I could ignore it, but since it's so strange, I am worried about it, and would like to at least know the cause... Thanks very much, Sompom Here are the source files hello-1.mod.c #include <linux/module.h> #include <linux/vermagic.h> #include <linux/compiler.h> MODULE_INFO(vermagic, VERMAGIC_STRING); struct module __this_module __attribute__((section(".gnu.linkonce.this_module"))) = { .name = KBUILD_MODNAME, .init = init_module, #ifdef CONFIG_MODULE_UNLOAD .exit = cleanup_module, #endif .arch = MODULE_ARCH_INIT, }; static const struct modversion_info ____versions[] __used __attribute__((section("__versions"))) = { { 0x3972220f, "module_layout" }, { 0xefd6cf06, "__aeabi_unwind_cpp_pr0" }, { 0xea147363, "printk" }, }; static const char __module_depends[] __used __attribute__((section(".modinfo"))) = "depends="; hello-1.c (modified slightly from the given link) /* hello-1.c - The simplest kernel module. * * Copyright (C) 2001 by Peter Jay Salzman * * 08/02/2006 - Updated by Rodrigo Rubira Branco <[email protected]> */ /* Kernel Programming */ #ifndef MODULE #define MODULE #endif #ifndef LINUX #define LINUX #endif #ifndef __KERNEL__ #define __KERNEL__ #endif #include <linux/module.h> /* Needed by all modules */ #include <linux/kernel.h> /* Needed for KERN_ALERT */ static int hello_init_module(void) { printk(KERN_ALERT "Hello world 1.\n"); /* A non 0 return means init_module failed; module can't be loaded.*/ return 0; } static void hello_cleanup_module(void) { printk(KERN_ALERT "Goodbye world 1.\n"); } module_init(hello_init_module); module_exit(hello_cleanup_module); MODULE_LICENSE("GPL"); Makefile export ARCH:=arm export CCPREFIX:=/opt/freescale/usr/local/gcc-4.4.4-glibc-2.11.1-multilib-1.0/arm-fsl-linux-gnueabi/bin/arm-linux- export CROSS_COMPILE:=${CCPREFIX} TARGET := hello-1 WARN := -W -Wall -Wstrict-prototypes -Wmissing-prototypes -Wno-sign-compare -Wno-unused -Werror UNUSED_FLAGS := -std=c99 -pedantic EXTRA_CFLAGS := -O2 -DMODULE -D__KERNEL__ ${WARN} ${INCLUDE} KDIR ?= /home/developer/src/ltib-microsys/ltib/rpm/BUILD/linux-2.6.35.3 ifneq ($(KERNELRELEASE),) # kbuild part of makefile obj-m := $(TARGET).o else # normal makefile default: clean $(MAKE) -C $(KDIR) M=$$PWD .PHONY: clean clean: -rm built-in.o -rm $(TARGET).ko -rm $(TARGET).ko.unsigned -rm $(TARGET).mod.c -rm $(TARGET).mod.o -rm $(TARGET).o -rm modules.order -rm Module.symvers endif

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  • Agile: User Stories for Machine Learning Project?

    - by benjismith
    I've just finished up with a prototype implementation of a supervised learning algorithm, automatically assigning categorical tags to all the items in our company database (roughly 5 million items). The results look good, and I've been given the go-ahead to plan the production implementation project. I've done this kind of work before, so I know how the functional components of the software. I need a collection of web crawlers to fetch data. I need to extract features from the crawled documents. Those documents need to be segregated into a "training set" and a "classification set", and feature-vectors need to be extracted from each document. Those feature vectors are self-organized into clusters, and the clusters are passed through a series of rebalancing operations. Etc etc etc etc. So I put together a plan, with about 30 unique development/deployment tasks, each with time estimates. The first stage of development -- ignoring some advanced features that we'd like to have in the long-term, but aren't high enough priority to make it into the development schedule yet -- is slated for about two months worth of work. (Keep in mind that I already have a working prototype, so the final implementation is significantly simpler than if the project was starting from scratch.) My manager said the plan looked good to him, but he asked if I could reorganize the tasks into user stories, for a few reasons: (1) our project management software is totally organized around user stories; (2) all of our scheduling is based on fitting entire user stories into sprints, rather than individually scheduling tasks; (3) other teams -- like the web developers -- have made great use of agile methodologies, and they've benefited from modelling all the software features as user stories. So I created a user story at the top level of the project: As a user of the system, I want to search for items by category, so that I can easily find the most relevant items within a huge, complex database. Or maybe a better top-level story for this feature would be: As a content editor, I want to automatically create categorical designations for the items in our database, so that customers can easily find high-value data within our huge, complex database. But that's not the real problem. The tricky part, for me, is figuring out how to create subordinate user stories for the rest of the machine learning architecture. Case in point... I know that the algorithm requires two major architectural subdivisions: (A) training, and (B) classification. And I know that the training portion of the architecture requires construction of a cluster-space. All the Agile Development literature I've read seems to indicate that a user story should be the "smallest possible implementation that provides any business value". And that makes a lot of sense when designing a piece of end-user software. Start small, and then incrementally add value when users demand additional functionality. But a cluster-space, in and of itself, provides zero business value. Nor does a crawler, or a feature-extractor. There's no business value (not for the end-user, or for any of the roles internal to the company) in a partial system. A trained cluster-space is only possible with the crawler and feature extractor, and only relevant if we also develop an accompanying classifier. I suppose it would be possible to create user stories where the subordinate components of the system act as the users in the stories: As a supervised-learning cluster-space construction routine, I want to consume data from a feature extractor, so that I can exist. But that seems really weird. What benefit does it provide me as the developer (or our users, or any other stakeholders, for that matter) to model my user stories like that? Although the main story can be easily divided along architectural-component boundaries (crawler, trainer, classifier, etc), I can't think of any useful decomposition from a user's perspective. What do you guys think? How do you plan Agile user stories for sophisticated, indivisible, non-user-facing components?

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  • Rails ActiveRecord friendly code from a Complex Join, Sum, and Group query

    - by Chad M
    PROBLEM Hello, I am having no luck trying to break down this SQL statement into ActiveRecord/Rails friendly code and I'd like to learn how I can avoid a find_by_sql statement in this situation. Scenario I have users that create audits when they perform an action. Each audit is of a specific audit_activity. Each audit_activity is worth a certain number of points, based on score_weight. I need to find the total scores of each user, based on their total accumulated audit_activity score_weights. Eventually I'll need to rank them which means adding a sort to this as well. My Code Here is my sql and simplified versions of the tables in question. Any thoughts? SQL with full column names (for clarity) SELECT users.id, u.email, SUM(audit_activity.score_weight) FROM users JOIN audits ON users.id = audits.user_id JOIN audit_activities ON audit_activities.id = audits.audit_activity_id GROUP BY users.id; Models: User, Audit, AuditActivity User fields: id, email class User < ActiveRecord::Base include Clearance::User has_many :audits end Audit fields: id, user_id, audit_activity_id class Audit < ActiveRecord::Base belongs_to :user belongs_to :audit_activity end AuditActivity fields: id, score_weight class AuditActivity < ActiveRecord::Base has_many :audits end Example Data Here is a set of SQL statements so you can play with similar data I'm working with and see what comes up when the concerned query is run. You should just be able to copy/paste the whole thing into a database query browser. CREATE TABLE users( id INTEGER NOT NULL, email TEXT (25), PRIMARY KEY (id) ); CREATE TABLE audits( id INTEGER NOT NULL, user_id INTEGER, audit_activity_id INTEGER, PRIMARY KEY (id) ); CREATE TABLE audit_activities( id INTEGER NOT NULL, score_weight INTEGER, PRIMARY KEY (id) ); INSERT INTO users(id, email) VALUES(1, "[email protected]"); INSERT INTO users(id, email) VALUES(2, "[email protected]"); INSERT INTO users(id, email) VALUES(3, "[email protected]"); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(1, 1, 1); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(2, 1, 2); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(3, 1, 1); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(4, 1, 3); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(5, 1, 1); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(6, 1, 4); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(7, 2, 4); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(8, 2, 4); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(9, 2, 4); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(10, 3, 3); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(11, 3, 2); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(12, 3, 2); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(13, 3, 2); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(14, 3, 3); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(15, 3, 1); INSERT INTO audits(id, user_id, audit_activity_id) VALUES(16, 3, 1); INSERT INTO audit_activities(id, score_weight) VALUES(1, 1); INSERT INTO audit_activities(id, score_weight) VALUES(2, 2); INSERT INTO audit_activities(id, score_weight) VALUES(3, 7); INSERT INTO audit_activities(id, score_weight) VALUES(4, 11); The Query Again, here is the query. SELECT u.id, u.email, SUM(aa.score_weight) FROM users u JOIN audits a ON u.id = a.user_id JOIN audit_activities aa ON aa.id = a.audit_activity_id GROUP BY u.id; Many Thanks, Chad

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  • L2S DataContext out of synch: row not found or changed

    - by awrigley
    The Problem I am getting a number of errors that imply that the DataContext, or rather, the way I am using the DataContext is getting out of synch. The error occurs on db.SubmitChanges() where db is my DataContext instance. The error is: Row not found or changed. The problem only occurs intermitently, for example, adding a row then deleting it. If I stop the dev server and restart, the added row is there and I can delete it no problem. Ie, it seems that the problem is related to the DataContext losing track of the rows that have been added. IMPORTANT: Before anyone votes to close this thread, on the basis of it being a duplicate, I have checked the sql server profiler and there is no "Where 0 = 1" in the SQL. I have also recreated the dbml file, so am satisfied that the database schema is in synch with the schema represented by the dbml file. Ie, no cases of mismatched nullable/not nullable columns, etc. My Diagnosis (for what it is worth): It seems to be a problem related to how I am using the DataContext. I am new to MVC, Repositories and Services patterns, so suspect that I have wired things up wrong. The Setup Simple eLearning app in its early stages. Pupils need to be able to add and delete courses (Courses table) to their UserCourses. To do this, I have a service that gets a specific DataContext instance Dependency Injected into its constructor. Service Class Constructor: public class SqlPupilBlockService : IPupilBlockService { DataContext db; public SqlPupilBlockService(DataContext db) { this.db = db; CoursesRepository = new SqlRepository<Course>(db); UserCoursesRepository = new SqlRepository<UserCourse>(db); } // Etc, etc } The CoursesRepository and UserCoursesRepository are both private properties of the service class that are of Type IRepository (just a simple generic repository interface). Sql Respoitory Code: public class SqlRepository<T> : IRepository<T> where T : class { DataContext db; public SqlRepository(DataContext db) { this.db = db; } #region IRepository<T> Members public IQueryable<T> Query { get { return db.GetTable<T>(); } } public List<T> FetchAll() { return Query.ToList(); } public void Add(T entity) { db.GetTable<T>().InsertOnSubmit(entity); } public void Delete(T entity) { db.GetTable<T>().DeleteOnSubmit(entity); } public void Save() { db.SubmitChanges(); } #endregion } The two methods for adding and deleting UserCourses are: Service Methods for Adding and Deleting UserCourses: public void AddUserCourse(int courseId) { UserCourse uc = new UserCourse(); uc.IdCourse = courseId; uc.IdUser = UserId; uc.DateCreated = DateTime.Now; uc.DateAmended = DateTime.Now; uc.Role = "Pupil"; uc.CourseNotes = string.Empty; uc.ActiveStepIndex = 0; UserCoursesRepository.Add(uc); UserCoursesRepository.Save(); } public void DeleteUserCourse(int courseId) { var uc = (UserCoursesRepository.Query.Where(x => x.IdUser == UserId && x.IdCourse == courseId)).Single(); UserCoursesRepository.Delete(uc); UserCoursesRepository.Save(); } Ajax I am using Ajax via Ajax.BeginForm I don't think that is relevant. ASP.NET MVC 3 I am using mvc3, but don't think that is relevant: the errors are related to model code.

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  • C++ Optimize if/else condition

    - by Heye
    I have a single line of code, that consumes 25% - 30% of the runtime of my application. It is a less-than comparator for an std::set (the set is implemented with a Red-Black-Tree). It is called about 180 Million times within 52 seconds. struct Entry { const float _cost; const long _id; // some other vars Entry(float cost, float id) : _cost(cost), _id(id) { } }; template<class T> struct lt_entry: public binary_function <T, T, bool> { bool operator()(const T &l, const T &r) const { // Most readable shape if(l._cost != r._cost) { return r._cost < l._cost; } else { return l._id < r._id; } } }; The entries should be sorted by cost and if the cost is the same by their id. I have many insertions for each extraction of the minimum. I thought about using Fibonacci-Heaps, but I have been told that they are theoretically nice, but suffer from high constants and are pretty complicated to implement. And since insert is in O(log(n)) the runtime increase is nearly constant with large n. So I think its okay to stick to the set. To improve performance I tried to express it in different shapes: return l._cost < r._cost || r._cost > l._cost || l._id < r._id; return l._cost < r._cost || (l._cost == r._cost && l._id < r._id); Even this: typedef union { float _f; int _i; } flint; //... flint diff; diff._f = (l._cost - r._cost); return (diff._i && diff._i >> 31) || l._id < r._id; But the compiler seems to be smart enough already, because I haven't been able to improve the runtime. I also thought about SSE but this problem is really not very applicable for SSE... The assembly looks somewhat like this: movss (%rbx),%xmm1 mov $0x1,%r8d movss 0x20(%rdx),%xmm0 ucomiss %xmm1,%xmm0 ja 0x410600 <_ZNSt8_Rb_tree[..]+96> ucomiss %xmm0,%xmm1 jp 0x4105fd <_ZNSt8_Rb_[..]_+93> jne 0x4105fd <_ZNSt8_Rb_[..]_+93> mov 0x28(%rdx),%rax cmp %rax,0x8(%rbx) jb 0x410600 <_ZNSt8_Rb_[..]_+96> xor %r8d,%r8d I have a very tiny bit experience with assembly language, but not really much. I thought it would be the best (only?) point to squeeze out some performance, but is it really worth the effort? Can you see any shortcuts that could save some cycles? The platform the code will run on is an ubuntu 12 with gcc 4.6 (-stl=c++0x) on a many-core intel machine. Only libraries available are boost, openmp and tbb. I am really stuck on this one, it seems so simple, but takes that much time. I have been crunching my head since days thinking how I could improve this line... Can you give me a suggestion how to improve this part, or is it already at its best?

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  • Windows Mobile 6.5 GPS Device - WaitForMultipleObjects returns 258 (timeout)

    - by wizmagister
    I’ve created a GPS program that track positions in realtime in the background for Windows mobile 6.1 in 2008-2009. It ran fine on these devices for many years. For some reason, the same code never worked perfectly on Windows Mobile 6.5. After many hour of operations (mostly when nobody use the device), I receive a “Timeout” (code 258) from the function "WaitForMultipleObjects": this.GPSEvent_WaitValue = WaitForMultipleObjects(2, this.GPSEvent_Handles, 0, 45000); Again, this can work for hours and suddenly, it's just impossible to get another position without : UPDATE: - Restarting the device (GoogleMap confirms that there's no GPS device present!) It has something to do with Windows Mobile going to sleep and slowing up my thread. Here's the core code (adapted from Microsoft SDK Sample): /// <summary> /// When "WindowsMobile" wake up the program to check for a new position /// </summary> private void OnNextGPSEvent_Callback() { int SecondsToNextWakeUp = ETL.Mobile.Device.ScheduledCallback.MINIMUM_SECONDTONEXTWAKEUP; switch (this.SleepingState) { case SleepingStateType.SleepingForNextPosition: // Get position this.GPSEvent_WaitValue = (WaitForEventThreadResultType)WaitForMultipleObjects(2, this.GPSEvent_Handles, 0, 45000); switch (this.GPSEvent_WaitValue) { case WaitForEventThreadResultType.Event_LocationChanged: // Got a new position this.FireLocationChanged(this.GetCurrentPosition()); // Manage device shutdown (save battery) if (this.PositionFrequency > MIN_SECONDS_FREQUENCY_FORDEVICE_SHUTDOWN) { // Close device this.CloseDevice(); SecondsToNextWakeUp = (this.PositionFrequency - GPSDEVICE_LOAD_SECONDS_LOAD_TIME); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; } else { // Default Wait Time this.SleepingState = SleepingStateType.SleepingForNextPosition; } break; case WaitForEventThreadResultType.Event_StateChanged: break; case WaitForEventThreadResultType.Timeout: case WaitForEventThreadResultType.Failed: case WaitForEventThreadResultType.Stop: // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // Too many errors this.ConsecutiveErrorReadingDevice++; if (this.ConsecutiveErrorReadingDevice > MAX_ERRORREADINGDEVICE) { this.CloseDevice(); SecondsToNextWakeUp = (this.PositionFrequency - GPSDEVICE_LOAD_SECONDS_LOAD_TIME); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; } else { // Default Wait Time this.SleepingState = SleepingStateType.SleepingForNextPosition; } break; } #endregion break; case SleepingStateType.SleepingBeforeDeviceWakeUp: this.OpenDevice(); SecondsToNextWakeUp = GPSDEVICE_LOAD_SECONDS_LOAD_TIME; this.SleepingState = SleepingStateType.SleepingForNextPosition; break; } if (this.IsListeningGPSEvent) { // Ajustement du prochain rappel this.NextGPSEvent_Callback.SecondToNextWakeUp = SecondsToNextWakeUp; this.NextGPSEvent_Callback.RequestWakeUpCallback(); } } /// <summary> ///Create Thread /// </summary> private void StartListeningThreadForGPSEvent() { // We only want to create the thread if we don't have one created already and we have opened the gps device if (this._GPSEventThread == null) { // Create and start thread to listen for GPS events this._GPSEventThread = new System.Threading.Thread(new System.Threading.ThreadStart(this.ListeningThreadForGPSEvent)); this._GPSEventThread.Start(); } } private void ListeningThreadForGPSEvent() { this.GPSEvent_WaitValue = WaitForEventThreadResultType.Stop; this.IsListeningGPSEvent = true; // Allocate handles worth of memory to pass to WaitForMultipleObjects this.GPSEvent_Handles = Helpers.LocalAlloc(12); Marshal.WriteInt32(this.GPSEvent_Handles, 0, this._StopHandle.ToInt32()); Marshal.WriteInt32(this.GPSEvent_Handles, 4, this._NewLocationHandle.ToInt32()); Marshal.WriteInt32(this.GPSEvent_Handles, 8, this._GPSDeviceStateChanged.ToInt32()); this.Start_NextGPSEvent_Timer(this.PositionFrequency); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; this.OnNextGPSEvent_Callback(); }

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  • How to select table column names in a view and pass to controller in rails?

    - by zachd1_618
    So I am new to Rails, and OO programming in general. I have some grasp of the MVC architecture. My goal is to make a (nearly) completely dynamic plug-and-play plotting web server. I am fairly confused with params, forms, and select helpers. What I want to do is use Rails drop downs to basically pass parameters as strings to my controller, which will use the params to select certain column data from my database and plot it dynamically. I have the latter part of the task working, but I can't seem to pass values from my view to controller. For simplicity's sake, say my database schema looks like this: --------------Plot--------------- |____x____|____y1____|____y2____| | 1 | 1 | 1 | | 2 | 2 | 4 | | 3 | 3 | 9 | | 4 | 4 | 16 | | 5 | 5 | 25 | ... and in my Model, I have dynamic selector scopes that will let me select just certain columns of data: in Plot.rb class Plot < ActiveRecord::Base scope :select_var, lambda {|varname| select(varname)} scope :between_x, lambda {|x1,x2| where("x BETWEEN ? and ?","#{x1}","#{x2}")} So this way, I can call: irb>>@p1 = Plot.select_var(['x','y1']).between_x(1,3) and get in return a class where @p1.x and @p1.y1 are my only attributes, only for values between x=1 to x=4, which I dynamically plot. I want to start off in a view (plot/index), where I can dynamically select which variable names (table column names), and which rows from the database to fetch and plot. The problem is, most select helpers don't seem to work with columns in the database, only rows. So to select columns, I first get an array of column names that exist in my database with a function I wrote. Plots Controller def index d=Plot.first @tags = d.list_vars end So @tags = ['x','y1','y2'] Then in my plot/index.html.erb I try to use a drop down to select wich variables I send back to the controller. index.html.erb <%= select_tag( :variable, options_for_select(@plots.first.list_vars,:name,:multiple=>:true) )%> <%= button_to 'Plot now!', :controller =>"plots/plot_vars", :variable => params[:variable]%> Finally, in the controller again Plots controller ... def plot_vars @plot_data=Plot.select_vars([params[:variable]]) end The problem is everytime I try this (or one of a hundred variations thereof), the params[:variable] is nill. How can I use a drop down to pass a parameter with string variable names to the controller? Sorry its so long, I have been struggling with this for about a month now. :-( I think my biggest problem is that this setup doesn't really match the Rails architecture. I don't have "users" and "articles" as individual entities. I really have a data structure, not a data object. Trying to work with the structure in terms of data object speak is not necessarily the easiest thing to do I think. For background: My actual database has about 250 columns and a couple million rows, and they get changed and modified from time to time. I know I can make the database smarter, but its not worth it on my end. I work at a scientific institute where there are a ton of projects with databases just like this. Each one has a web developer that spends months setting up a web interface and their own janky plotting setups. I want to make this completely dynamic, as a plug-and-play solution so all you have to do is specify your database connection, and this rails setup will automatically show and plot which data you want in it. I am more of a sequential programmer and number cruncher, as are many people here. I think this project could be very helpful in the end, but its difficult to figure out for me right now.

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  • CakePHP access indirectly related model - beginner's question

    - by user325077
    Hi everyone, I am writing a CakePHP application to log the work I do for various clients, but after trying for days I seem unable to get it to do what I want. I have read most of the book CakePHP's website. and googled for all I'm worth, so I presume I am missing something obvious! Every 'log item' belongs to a 'sub-project, which in turn belongs to a 'project', which in turn belongs to a 'sub-client' which finally belongs to a client. These are the 5 MySQL tables I am using: mysql> DESCRIBE log_items; +-----------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +-----------------+--------------+------+-----+---------+----------------+ | id | int(11) | NO | PRI | NULL | auto_increment | | date | date | NO | | NULL | | | time | time | NO | | NULL | | | time_spent | int(11) | NO | | NULL | | | sub_projects_id | int(11) | NO | MUL | NULL | | | title | varchar(100) | NO | | NULL | | | description | text | YES | | NULL | | | created | datetime | YES | | NULL | | | modified | datetime | YES | | NULL | | +-----------------+--------------+------+-----+---------+----------------+ mysql> DESCRIBE sub_projects; +-------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +-------------+--------------+------+-----+---------+----------------+ | id | int(11) | NO | PRI | NULL | auto_increment | | name | varchar(100) | NO | | NULL | | | projects_id | int(11) | NO | MUL | NULL | | | created | datetime | YES | | NULL | | | modified | datetime | YES | | NULL | | +-------------+--------------+------+-----+---------+----------------+ mysql> DESCRIBE projects; +----------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +----------------+--------------+------+-----+---------+----------------+ | id | int(11) | NO | PRI | NULL | auto_increment | | name | varchar(100) | NO | | NULL | | | sub_clients_id | int(11) | NO | MUL | NULL | | | created | datetime | YES | | NULL | | | modified | datetime | YES | | NULL | | +----------------+--------------+------+-----+---------+----------------+ mysql> DESCRIBE sub_clients; +------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +------------+--------------+------+-----+---------+----------------+ | id | int(11) | NO | PRI | NULL | auto_increment | | name | varchar(100) | NO | | NULL | | | clients_id | int(11) | NO | MUL | NULL | | | created | datetime | YES | | NULL | | | modified | datetime | YES | | NULL | | +------------+--------------+------+-----+---------+----------------+ mysql> DESCRIBE clients; +----------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +----------+--------------+------+-----+---------+----------------+ | id | int(11) | NO | PRI | NULL | auto_increment | | name | varchar(100) | NO | | NULL | | | created | datetime | YES | | NULL | | | modified | datetime | YES | | NULL | | +----------+--------------+------+-----+---------+----------------+ I have set up the following associations in CakePHP: LogItem belongsTo SubProjects SubProject belongsTo Projects Project belongsTo SubClients SubClient belongsTo Clients Client hasMany SubClients SubClient hasMany Projects Project hasMany SubProjects SubProject hasMany LogItems Using 'cake bake' I have created the models, controllers (index, view add, edit and delete) and views, and things seem to function - as in I am able to perform simple CRUD operations successfully. The Question When editing a 'log item' at www.mydomain/log_items/edit I am presented with the view you would all suspect; namely the columns of the log_items table with the appropriate textfields/select boxes etc. I would also like to incorporate select boxes to choose the client, sub-client, project and sub-project in the 'log_items' edit view. Ideally the 'sub-client' select box should populate itself depending upon the 'client' chosen, the 'project' select box should also populate itself depending on the 'sub-client' selected etc, etc. I guess the way to go about populating the select boxes with relevant options is Ajax, but I am unsure of how to go about actually accessing a model from the child view of a indirectly related model, for example how to create a 'sub-client' select box in the 'log_items' edit view. I have have found this example: http://forum.phpsitesolutions.com/php-frameworks/cakephp/ajax-cakephp-dynamically-populate-html-select-dropdown-box-t29.html where someone achieves something similar for US states, counties and cities. However, I noticed in the database schema - which is downloadable from the site above link - that the database tables don't have any foreign keys, so now I'm wondering if I'm going about things in the correct manner. Any pointers and advice would be very much appreciated. Kind regards, Chris

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  • Upgrading from TFS 2010 RC to TFS 2010 RTM done

    - by Martin Hinshelwood
    Today is the big day, with the Launch of Visual Studio 2010 already done in Asia, and rolling around the world towards us, we are getting ready for the RTM (Released). We have had TFS 2010 in Production for nearly 6 months and have had only minimal problems. Update 12th April 2010  – Added Scott Hanselman’s tweet about the MSDN download release time. SSW was the first company in the world outside of Microsoft to deploy Visual Studio 2010 Team Foundation Server to production, not once, but twice. I am hoping to make it 3 in a row, but with all the hype around the new version, and with it being a production release and not just a go-live, I think there will be a lot of competition. Developers: MSDN will be updated with #vs2010 downloads and details at 10am PST *today*! @shanselman - Scott Hanselman Same as before, we need to Uninstall 2010 RC and install 2010 RTM. The installer will take care of all the complexity of actually upgrading any schema changes. If you are upgrading from TFS 2008 to TFS2010 you can follow our Rules To Better TFS 2010 Migration and read my post on our successes.   We run TFS 2010 in a Hyper-V virtual environment, so we have the advantage of running a snapshot as well as taking a DB backup. Done - Snapshot the hyper-v server Microsoft does not support taking a snapshot of a running server, for very good reason, and Brian Harry wrote a post after my last upgrade with the reason why you should never snapshot a running server. Done - Uninstall Visual Studio Team Explorer 2010 RC You will need to uninstall all of the Visual Studio 2010 RC client bits that you have on the server. Done - Uninstall TFS 2010 RC Done - Install TFS 2010 RTM Done - Configure TFS 2010 RTM Pick the Upgrade option and point it at your existing “tfs_Configuration” database to load all of the existing settings Done - Upgrade the SharePoint Extensions Upgrade Build Servers (Pending) Test the server The back out plan, and you should always have one, is to restore the snapshot. Upgrading to Team Foundation Server 2010 – Done The first thing you need to do is off the TFS server and then log into the Hyper-v server and create a snapshot. Figure: Make sure you turn the server off and delete all old snapshots before you take a new one I noticed that the snapshot that was taken before the Beta 2 to RC upgrade was still there. You should really delete old snapshots before you create a new one, but in this case the SysAdmin (who is currently tucked up in bed) asked me not to. I guess he is worried about a developer messing up his server Turn your server on and wait for it to boot in anticipation of all the nice shiny RTM’ness that is coming next. The upgrade procedure for TFS2010 is to uninstal the old version and install the new one. Figure: Remove Visual Studio 2010 Team Foundation Server RC from the system.   Figure: Most of the heavy lifting is done by the Uninstaller, but make sure you have removed any of the client bits first. Specifically Visual Studio 2010 or Team Explorer 2010.  Once the uninstall is complete, this took around 5 minutes for me, you can begin the install of the RTM. Running the 64 bit OS will allow the application to use more than 2GB RAM, which while not common may be of use in heavy load situations. Figure: It is always recommended to install the 64bit version of a server application where possible. I do not think it is likely, with SharePoint 2010 and Exchange 2010  and even Windows Server 2008 R2 being 64 bit only, I do not think there will be another release of a server app that is 32bit. You then need to choose what it is you want to install. This depends on how you are running TFS and on how many servers. In our case we run TFS and the Team Foundation Build Service (controller only) on out TFS server along with Analysis services and Reporting Services. But our SharePoint server lives elsewhere. Figure: This always confuses people, but in reality it makes sense. Don’t install what you do not need. Every extra you install has an impact of performance. If you are integrating with SharePoint you will need to run this install on every Front end server in your farm and don’t forget to upgrade your Build servers and proxy servers later. Figure: Selecting only Team Foundation Server (TFS) and Team Foundation Build Services (TFBS)   It is worth noting that if you have a lot of builds kicking off, and hence a lot of get operations against your TFS server, you can use a proxy server to cache the source control on another server in between your TFS server and your build servers. Figure: Installing Microsoft .NET Framework 4 takes the most time. Figure: Now run Windows Update, and SSW Diagnostic to make sure all your bits and bobs are up to date. Note: SSW Diagnostic will check your Power Tools, Add-on’s, Check in Policies and other bits as well. Configure Team Foundation Server 2010 – Done Now you can configure the server. If you have no key you will need to pick “Install a Trial Licence”, but it is only £500, or free with a MSDN subscription. Anyway, if you pick Trial you get 90 days to get your key. Figure: You can pick trial and add your key later using the TFS Server Admin. Here is where the real choices happen. We are doing an Upgrade from a previous version, so I will pick Upgrade the same as all you folks that are using the RC or TFS 2008. Figure: The upgrade wizard takes your existing 2010 or 2008 databases and upgraded them to the release.   Once you have entered your database server name you can click “List available databases” and it will show what it can upgrade. Figure: Select your database from the list and at this point, make sure you have a valid backup. At this point you have not made ANY changes to the databases. At this point the configuration wizard will load configuration from your existing database if you have one. If you are upgrading TFS 2008 refer to Rules To Better TFS 2010 Migration. Mostly during the wizard the default values will suffice, but depending on the configuration you want you can pick different options. Figure: Set the application tier account and Authentication method to use. We use NTLM to keep things simple as we host our TFS server externally for our remote developers.  Figure: Setting your TFS server URL’s to be the remote URL’s allows the reports to be accessed without using VPN. Very handy for those remote developers. Figure: Detected the existing Warehouse no problem. Figure: Again we love green ticks. It gives us a warm fuzzy feeling. Figure: The username for connecting to Reporting services should be a domain account (if you are on a domain that is). Figure: Setup the SharePoint integration to connect to your external SharePoint server. You can take the option to connect later.   You then need to run all of your readiness checks. These check can save your life! it will check all of the settings that you have entered as well as checking all the external services are configures and running properly. There are two reasons that TFS 2010 is so easy and painless to install where previous version were not. Microsoft changes the install to two steps, Install and configuration. The second reason is that they have pulled out all of the stops in making the install run all the checks necessary to make sure that once you start the install that it will complete. if you find any errors I recommend that you report them on http://connect.microsoft.com so everyone can benefit from your misery.   Figure: Now we have everything setup the configuration wizard can do its work.  Figure: Took a while on the “Web site” stage for some point, but zipped though after that.  Figure: last wee bit. TFS Needs to do a little tinkering with the data to complete the upgrade. Figure: All upgraded. I am not worried about the yellow triangle as SharePoint was being a little silly Exception Message: TF254021: The account name or password that you specified is not valid. (type TfsAdminException) Exception Stack Trace:    at Microsoft.TeamFoundation.Management.Controls.WizardCommon.AccountSelectionControl.TestLogon(String connectionString)    at System.ComponentModel.BackgroundWorker.WorkerThreadStart(Object argument) [Info   @16:10:16.307] Benign exception caught as part of verify: Exception Message: TF255329: The following site could not be accessed: http://projects.ssw.com.au/. The server that you specified did not return the expected response. Either you have not installed the Team Foundation Server Extensions for SharePoint Products on this server, or a firewall is blocking access to the specified site or the SharePoint Central Administration site. For more information, see the Microsoft Web site (http://go.microsoft.com/fwlink/?LinkId=161206). (type TeamFoundationServerException) Exception Stack Trace:    at Microsoft.TeamFoundation.Client.SharePoint.WssUtilities.VerifyTeamFoundationSharePointExtensions(ICredentials credentials, Uri url)    at Microsoft.TeamFoundation.Admin.VerifySharePointSitesUrl.Verify() Inner Exception Details: Exception Message: TF249064: The following Web service returned an response that is not valid: http://projects.ssw.com.au/_vti_bin/TeamFoundationIntegrationService.asmx. This Web service is used for the Team Foundation Server Extensions for SharePoint Products. Either the extensions are not installed, the request resulted in HTML being returned, or there is a problem with the URL. Verify that the following URL points to a valid SharePoint Web application and that the application is available: http://projects.ssw.com.au. If the URL is correct and the Web application is operating normally, verify that a firewall is not blocking access to the Web application. (type TeamFoundationServerInvalidResponseException) Exception Data Dictionary: ResponseStatusCode = InternalServerError I’ll look at SharePoint after, probably the SharePoint box just needs a restart or a kick If there is a problem with SharePoint it will come out in testing, But I will definatly be passing this on to Microsoft.   Upgrading the SharePoint connector to TFS 2010 You will need to upgrade the Extensions for SharePoint Products and Technologies on all of your SharePoint farm front end servers. To do this uninstall  the TFS 2010 RC from it in the same way as the server, and then install just the RTM Extensions. Figure: Only install the SharePoint Extensions on your SharePoint front end servers. TFS 2010 supports both SharePoint 2007 and SharePoint 2010.   Figure: When you configure SharePoint it uploads all of the solutions and templates. Figure: Everything is uploaded Successfully. Figure: TFS even remembered the settings from the previous installation, fantastic.   Upgrading the Team Foundation Build Servers to TFS 2010 Just like on the SharePoint servers you will need to upgrade the Build Server to the RTM. Just uninstall TFS 2010 RC and then install only the Team Foundation Build Services component. Unlike on the SharePoint server you will probably have some version of Visual Studio installed. You will need to remove this as well. (Coming Soon) Connecting Visual Studio 2010 / 2008 / 2005 and Eclipse to TFS2010 If you have developers still on Visual Studio 2005 or 2008 you will need do download the respective compatibility pack: Visual Studio Team System 2005 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 Visual Studio Team System 2008 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 If you are using Eclipse you can download the new Team Explorer Everywhere install for connecting to TFS. Get your developers to check that you have the latest version of your applications with SSW Diagnostic which will check for Service Packs and hot fixes to Visual Studio as well.   Technorati Tags: TFS,TFS2010,TFS 2010,Upgrade

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  • SQL SERVER – Shrinking NDF and MDF Files – Readers’ Opinion

    - by pinaldave
    Previously, I had written a blog post about SQL SERVER – Shrinking NDF and MDF Files – A Safe Operation. After that, I have written the following blog post that talks about the advantage and disadvantage of Shrinking and why one should not be Shrinking a file SQL SERVER – SHRINKFILE and TRUNCATE Log File in SQL Server 2008. On this subject, SQL Server Expert Imran Mohammed left an excellent comment. I just feel that his comment is worth a big article itself. For everybody to read his wonderful explanation, I am posting this blog post here. Thanks Imran! Shrinking Database always creates performance degradation and increases fragmentation in the database. I suggest that you keep that in mind before you start reading the following comment. If you are going to say Shrinking Database is bad and evil, here I am saying it first and loud. Now, the comment of Imran is written while keeping in mind only the process showing how the Shrinking Database Operation works. Imran has already explained his understanding and requests further explanation. I have removed the Best Practices section from Imran’s comments, as there are a few corrections. Comments from Imran - Before I explain to you the concept of Shrink Database, let us understand the concept of Database Files. When we create a new database inside the SQL Server, it is typical that SQl Server creates two physical files in the Operating System: one with .MDF Extension, and another with .LDF Extension. .MDF is called as Primary Data File. .LDF is called as Transactional Log file. If you add one or more data files to a database, the physical file that will be created in the Operating System will have an extension of .NDF, which is called as Secondary Data File; whereas, when you add one or more log files to a database, the physical file that will be created in the Operating System will have the same extension as .LDF. The questions now are, “Why does a new data file have a different extension (.NDF)?”, “Why is it called as a secondary data file?” and, “Why is .MDF file called as a primary data file?” Answers: Note: The following explanation is based on my limited knowledge of SQL Server, so experts please do comment. A data file with a .MDF extension is called a Primary Data File, and the reason behind it is that it contains Database Catalogs. Catalogs mean Meta Data. Meta Data is “Data about Data”. An example for Meta Data includes system objects that store information about other objects, except the data stored by the users. sysobjects stores information about all objects in that database. sysindexes stores information about all indexes and rows of every table in that database. syscolumns stores information about all columns that each table has in that database. sysusers stores how many users that database has. Although Meta Data stores information about other objects, it is not the transactional data that a user enters; rather, it’s a system data about the data. Because Primary Data File (.MDF) contains important information about the database, it is treated as a special file. It is given the name Primary Data file because it contains the Database Catalogs. This file is present in the Primary File Group. You can always create additional objects (Tables, indexes etc.) in the Primary data file (This file is present in the Primary File group), by mentioning that you want to create this object under the Primary File Group. Any additional data file that you add to the database will have only transactional data but no Meta Data, so that’s why it is called as the Secondary Data File. It is given the extension name .NDF so that the user can easily identify whether a specific data file is a Primary Data File or a Secondary Data File(s). There are many advantages of storing data in different files that are under different file groups. You can put your read only in the tables in one file (file group) and read-write tables in another file (file group) and take a backup of only the file group that has read the write data, so that you can avoid taking the backup of a read-only data that cannot be altered. Creating additional files in different physical hard disks also improves I/O performance. A real-time scenario where we use Files could be this one: Let’s say you have created a database called MYDB in the D-Drive which has a 50 GB space. You also have 1 Database File (.MDF) and 1 Log File on D-Drive and suppose that all of that 50 GB space has been used up and you do not have any free space left but you still want to add an additional space to the database. One easy option would be to add one more physical hard disk to the server, add new data file to MYDB database and create this new data file in a new hard disk then move some of the objects from one file to another, and put the file group under which you added new file as default File group, so that any new object that is created gets into the new files, unless specified. Now that we got a basic idea of what data files are, what type of data they store and why they are named the way they are, let’s move on to the next topic, Shrinking. First of all, I disagree with the Microsoft terminology for naming this feature as “Shrinking”. Shrinking, in regular terms, means to reduce the size of a file by means of compressing it. BUT in SQL Server, Shrinking DOES NOT mean compressing. Shrinking in SQL Server means to remove an empty space from database files and release the empty space either to the Operating System or to SQL Server. Let’s examine this through an example. Let’s say you have a database “MYDB” with a size of 50 GB that has a free space of about 20 GB, which means 30GB in the database is filled with data and the 20 GB of space is free in the database because it is not currently utilized by the SQL Server (Database); it is reserved and not yet in use. If you choose to shrink the database and to release an empty space to Operating System, and MIND YOU, you can only shrink the database size to 30 GB (in our example). You cannot shrink the database to a size less than what is filled with data. So, if you have a database that is full and has no empty space in the data file and log file (you don’t have an extra disk space to set Auto growth option ON), YOU CANNOT issue the SHRINK Database/File command, because of two reasons: There is no empty space to be released because the Shrink command does not compress the database; it only removes the empty space from the database files and there is no empty space. Remember, the Shrink command is a logged operation. When we perform the Shrink operation, this information is logged in the log file. If there is no empty space in the log file, SQL Server cannot write to the log file and you cannot shrink a database. Now answering your questions: (1) Q: What are the USEDPAGES & ESTIMATEDPAGES that appear on the Results Pane after using the DBCC SHRINKDATABASE (NorthWind, 10) ? A: According to Books Online (For SQL Server 2000): UsedPages: the number of 8-KB pages currently used by the file. EstimatedPages: the number of 8-KB pages that SQL Server estimates the file could be shrunk down to. Important Note: Before asking any question, make sure you go through Books Online or search on the Google once. The reasons for doing so have many advantages: 1. If someone else already has had this question before, chances that it is already answered are more than 50 %. 2. This reduces your waiting time for the answer. (2) Q: What is the difference between Shrinking the Database using DBCC command like the one above & shrinking it from the Enterprise Manager Console by Right-Clicking the database, going to TASKS & then selecting SHRINK Option, on a SQL Server 2000 environment? A: As far as my knowledge goes, there is no difference, both will work the same way, one advantage of using this command from query analyzer is, your console won’t be freezed. You can do perform your regular activities using Enterprise Manager. (3) Q: What is this .NDF file that is discussed above? I have never heard of it. What is it used for? Is it used by end-users, DBAs or the SERVER/SYSTEM itself? A: .NDF File is a secondary data file. You never heard of it because when database is created, SQL Server creates database by default with only 1 data file (.MDF) and 1 log file (.LDF) or however your model database has been setup, because a model database is a template used every time you create a new database using the CREATE DATABASE Command. Unless you have added an extra data file, you will not see it. This file is used by the SQL Server to store data which are saved by the users. Hope this information helps. I would like to as the experts to please comment if what I understand is not what the Microsoft guys meant. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Readers Contribution, Readers Question, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Running ASP.NET Webforms and ASP.NET MVC side by side

    - by rajbk
    One of the nice things about ASP.NET MVC and its older brother ASP.NET WebForms is that they are both built on top of the ASP.NET runtime environment. The advantage of this is that, you can still run them side by side even though MVC and WebForms are different frameworks. Another point to note is that with the release of the ASP.NET routing in .NET 3.5 SP1, we are able to create SEO friendly URLs that do not map to specific files on disk. The routing is part of the core runtime environment and therefore can be used by both WebForms and MVC. To run both frameworks side by side, we could easily create a separate folder in your MVC project for all our WebForm files and be good to go. What this post shows you instead, is how to have an MVC application with WebForm pages  that both use a common master page and common routing for SEO friendly URLs.  A sample project that shows WebForms and MVC running side by side is attached at the bottom of this post. So why would we want to run WebForms and MVC in the same project?  WebForms come with a lot of nice server controls that provide a lot of functionality. One example is the ReportViewer control. Using this control and client report definition files (RDLC), we can create rich interactive reports (with charting controls). I show you how to use the ReportViewer control in a WebForm project here :  Creating an ASP.NET report using Visual Studio 2010. We can create even more advanced reports by using SQL reporting services that can also be rendered by the ReportViewer control. Now, consider the sample MVC application I blogged about called ASP.NET MVC Paging/Sorting/Filtering using the MVCContrib Grid and Pager. Assume you were given the requirement to add a UI to the MVC application where users could interact with a report and be given the option to export the report to Excel, PDF or Word. How do you go about doing it?   This is a perfect scenario to use the ReportViewer control and RDLCs. As you saw in the post on creating the ASP.NET report, the ReportViewer control is a Web Control and is designed to be run in a WebForm project with dependencies on, amongst others, a ScriptManager control and the beloved Viewstate.  Since MVC and WebForm both run under the same runtime, the easiest thing to is to add the WebForm application files (index.aspx, rdlc, related class files) into our MVC project. You can copy the files over from the WebForm project into the MVC project. Create a new folder in our MVC application called CommonReports. Add the index.aspx and rdlc file from the Webform project   Right click on the Index.aspx file and convert it to a web application. This will add the index.aspx.designer.cs file (this step is not required if you are manually adding a WebForm aspx file into the MVC project).    Verify that all the type names for the ObjectDataSources in code behind to point to the correct ProductRepository and fix any compiler errors. Right click on Index.aspx and select “View in browser”. You should see a screen like the one below:   There are two issues with our page. It does not use our site master page and the URL is not SEO friendly. Common Master Page The easiest way to use master pages with both MVC and WebForm pages is to have a common master page that each inherits from as shown below. The reason for this is most WebForm controls require them to be inside a Form control and require ControlState or ViewState. ViewMasterPages used in MVC, on the other hand, are designed to be used with content pages that derive from ViewPage with Viewstate turned off. By having a separate master page for MVC and WebForm that inherit from the Root master page,, we can set properties that are specific to each. For example, in the Webform master, we can turn on ViewState, add a form tag etc. Another point worth noting is that if you set a WebForm page to use a MVC site master page, you may run into errors like the following: A ViewMasterPage can be used only with content pages that derive from ViewPage or ViewPage<TViewItem> or Control 'MainContent_MyButton' of type 'Button' must be placed inside a form tag with runat=server. Since the ViewMasterPage inherits from MasterPage as seen below, we make our Root.master inherit from MasterPage, MVC.master inherit from ViewMasterPage and Webform.master inherits from MasterPage. We define the attributes on the master pages like so: Root.master <%@ Master Inherits="System.Web.UI.MasterPage"  … %> MVC.master <%@ Master MasterPageFile="~/Views/Shared/Root.Master" Inherits="System.Web.Mvc.ViewMasterPage" … %> WebForm.master <%@ Master MasterPageFile="~/Views/Shared/Root.Master" Inherits="NorthwindSales.Views.Shared.Webform" %> Code behind: public partial class Webform : System.Web.UI.MasterPage {} We make changes to our reports aspx file to use the Webform.master. See the source of the master pages in the sample project for a better understanding of how they are connected. SEO friendly links We want to create SEO friendly links that point to our report. A request to /Reports/Products should render the report located in ~/CommonReports/Products.aspx. Simillarly to support future reports, a request to /Reports/Sales should render a report in ~/CommonReports/Sales.aspx. Lets start by renaming our index.aspx file to Products.aspx to be consistent with our routing criteria above. As mentioned earlier, since routing is part of the core runtime environment, we ca easily create a custom route for our reports by adding an entry in Global.asax. public static void RegisterRoutes(RouteCollection routes) { routes.IgnoreRoute("{resource}.axd/{*pathInfo}");   //Custom route for reports routes.MapPageRoute( "ReportRoute", // Route name "Reports/{reportname}", // URL "~/CommonReports/{reportname}.aspx" // File );     routes.MapRoute( "Default", // Route name "{controller}/{action}/{id}", // URL with parameters new { controller = "Home", action = "Index", id = UrlParameter.Optional } // Parameter defaults ); } With our custom route in place, a request to Reports/Employees will render the page at ~/CommonReports/Employees.aspx. We make this custom route the first entry since the routing system walks the table from top to bottom, and the first route to match wins. Note that it is highly recommended that you write unit tests for your routes to ensure that the mappings you defined are correct. Common Menu Structure The master page in our original MVC project had a menu structure like so: <ul id="menu"> <li> <%=Html.ActionLink("Home", "Index", "Home") %></li> <li> <%=Html.ActionLink("Products", "Index", "Products") %></li> <li> <%=Html.ActionLink("Help", "Help", "Home") %></li> </ul> We want this menu structure to be common to all pages/views and hence should reside in Root.master. Unfortunately the Html.ActionLink helpers will not work since Root.master inherits from MasterPage which does not have the helper methods available. The quickest way to resolve this issue is to use RouteUrl expressions. Using  RouteUrl expressions, we can programmatically generate URLs that are based on route definitions. By specifying parameter values and a route name if required, we get back a URL string that corresponds to a matching route. We move our menu structure to Root.master and change it to use RouteUrl expressions: <ul id="menu"> <li> <asp:HyperLink ID="hypHome" runat="server" NavigateUrl="<%$RouteUrl:routename=default,controller=home,action=index%>">Home</asp:HyperLink></li> <li> <asp:HyperLink ID="hypProducts" runat="server" NavigateUrl="<%$RouteUrl:routename=default,controller=products,action=index%>">Products</asp:HyperLink></li> <li> <asp:HyperLink ID="hypReport" runat="server" NavigateUrl="<%$RouteUrl:routename=ReportRoute,reportname=products%>">Product Report</asp:HyperLink></li> <li> <asp:HyperLink ID="hypHelp" runat="server" NavigateUrl="<%$RouteUrl:routename=default,controller=home,action=help%>">Help</asp:HyperLink></li> </ul> We are done adding the common navigation to our application. The application now uses a common theme, routing and navigation structure. Conclusion We have seen how to do the following through this post Add a WebForm page from a WebForm project to an existing ASP.NET MVC application Use a common master page for both WebForm and MVC pages Use routing for SEO friendly links Use a common menu structure for both WebForm and MVC. The sample project is attached below. Version: VS 2010 RTM Remember to change your connection string to point to your Northwind database NorthwindSalesMVCWebform.zip

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  • Upgrading Team Foundation Server 2008 to 2010

    - by Martin Hinshelwood
    I am sure you will have seen my posts on upgrading our internal Team Foundation Server from TFS2008 to TFS2010 Beta 2, RC and RTM, but what about a fresh upgrade of TFS2008 to TFS2010 using the RTM version of TFS. One of our clients is taking the plunge with TFS2010, so I have the job of doing the upgrade. It is sometimes very useful to have a team member that starts work when most of the Sydney workers are heading home as I can do the upgrade without impacting them. The down side is that if you have any blockers then you can be pretty sure that everyone that can deal with your problem is asleep I am starting with an existing blank installation of TFS 2010, but Adam Cogan let slip that he was the one that did the install so I thought it prudent to make sure that it was OK. Verifying Team Foundation Server 2010 We need to check that TFS 2010 has been installed correctly. First, check the Admin console and have a root about for any errors. Figure: Even the SQL Setup looks good. I don’t know how Adam did it! Backing up the Team Foundation Server 2008 Databases As we are moving from one server to another (recommended method) we will be taking a backup of our TFS2008 databases and resorting them to the SQL Server for the new TFS2010 Server. Do not just detach and reattach. This will cause problems with the version of the database. If you are running a test migration you just need to create a backup of the TFS 2008 databases, but if you are doing the live migration then you should stop IIS on the TFS 2008 server before you backup the databases. This will stop any inadvertent check-ins or changes to TFS 2008. Figure: Stop IIS before you take a backup to prevent any TFS 2008 changes being written to the database. It is good to leave a little time between taking the TFS 2008 server offline and commencing the upgrade as there is always one developer who has not finished and starts screaming. This time it was John Liu that needed 10 more minutes to make his changes and check-in, so I always give it 30 minutes and see if anyone screams. John Liu [SSW] said:   are you doing something to TFS :-O MrHinsh [SSW UK][VS ALM MVP] said:   I have stopped TFS 2008 as per my emails John Liu [SSW] said:   haven't finish check in @_@   can we have it for 10mins? :) MrHinsh [SSW UK][VS ALM MVP] said:   TFS 2008 has been started John Liu [SSW] said:   I love you! -IM conversation at TFS Upgrade +25 minutes After John confirmed that he had everything done I turned IIS off again and made a cup of tea. There were no more screams so the upgrade can continue. Figure: Backup all of the databases for TFS and include the Reporting Services, just in case.   Figure: Check that all the backups have been taken Once you have your backups, you need to copy them to your new TFS2010 server and restore them. This is a good way to proceed as if we have any problems, or just plain run out of time, then you just turn the TFS 2008 server back on and all you have lost is one upgrade day, and not 10 developer days. As per the rules, you should record the number of files and the total number of areas and iterations before the upgrade so you have something to compare to: TFS2008 File count: Type Count 1 1845 2 15770 Areas & Iterations: 139 You can use this to verify that the upgrade was successful. it should however be noted that the numbers in TFS 2010 will be bigger. This is due to some of the sorting out that TFS does during the upgrade process. Restore Team Foundation Server 2008 Databases Restoring the databases is much more time consuming than just attaching them as you need to do them one at a time. But you may be taking a backup of an operational database and need to restore all your databases to a particular point in time instead of to the latest. I am doing latest unless I encounter any problems. Figure: Restore each of the databases to either a latest or specific point in time.     Figure: Restore all of the required databases Now that all of your databases are restored you now need to upgrade them to Team Foundation Server 2010. Upgrade Team Foundation Server 2008 Databases This is probably the easiest part of the process. You need to call a fire and forget command that will go off to the database specified, find the TFS 2008 databases and upgrade them to 2010. During this process all of the 6 main TFS 2008 databases are merged into the TfsVersionControl database, upgraded and then the database is renamed to TFS_[CollectionName]. The rename is only the database and not the physical files, so it is worth going back and renaming the physical file as well. This keeps everything neat and tidy. If you plan to keep the old TFS 2008 server around, for example if you are doing a test migration first, then you will need to change the TFS GUID. This GUID is unique to each TFS instance and is preserved when you upgrade. This GUID is used by the clients and they can get a little confused if there are two servers with the same one. To kick of the upgrade you need to open a command prompt and change the path to “C:\Program Files\Microsoft Team Foundation Server 2010\Tools” and run the “import” command in  “tfsconfig”. TfsConfig import /sqlinstance:<Previous TFS Data Tier>                  /collectionName:<Collection Name>                  /confirmed Imports a TFS 2005 or 2008 data tier as a new project collection. Important: This command should only be executed after adequate backups have been performed. After you import, you will need to configure portal and reporting settings via the administration console. EXAMPLES -------- TfsConfig import /sqlinstance:tfs2008sql /collectionName:imported /confirmed TfsConfig import /sqlinstance:tfs2008sql\Instance /collectionName:imported /confirmed OPTIONS: -------- sqlinstance         The sql instance of the TFS 2005 or 2008 data tier. The TFS databases at that location will be modified directly and will no longer be usable as previous version databases.  Ensure you have back-ups. collectionName      The name of the new Team Project Collection. confirmed           Confirm that you have backed-up databases before importing. This command will automatically look for the TfsIntegration database and verify that all the other required databases exist. In this case it took around 5 minutes to complete the upgrade as the total database size was under 700MB. This was unlike the upgrade of SSW’s production database with over 17GB of data which took a few hours. At the end of the process you should get no errors and no warnings. The Upgrade operation on the ApplicationTier feature has completed. There were 0 errors and 0 warnings. As this is a new server and not a pure upgrade there should not be a problem with the GUID. If you think at any point you will be doing this more than once, for example doing a test migration, or merging many TFS 2008 instances into a single one, then you should go back and rename the physical TfsVersionControl.mdf file to the same as the new collection. This will avoid confusion later down the line. To do this, detach the new collection from the server and rename the physical files. Then reattach and change the physical file locations to match the new name. You can follow http://www.mssqltips.com/tip.asp?tip=1122 for a more detailed explanation of how to do this. Figure: Stop the collection so TFS does not take a wobbly when we detach the database. When you try to start the new collection again you will get a conflict with project names and will require to remove the Test Upgrade collection. This is fine and it just needs detached. Figure: Detaching the test upgrade from the new Team Foundation Server 2010 so we can start the new Collection again. You will now be able to start the new upgraded collection and you are ready for testing. Do you remember the stats we took off the TFS 2008 server? TFS2008 File count: Type Count 1 1845 2 15770 Areas & Iterations: 139 Well, now we need to compare them to the TFS 2010 stats, remembering that there will probably be more files under source control. TFS2010 File count: Type Count 1 19288 Areas & Iterations: 139 Lovely, the number of iterations are the same, and the number of files is bigger. Just what we were looking for. Testing the upgraded Team Foundation Server 2010 Project Collection Can we connect to the new collection and project? Figure: We can connect to the new collection and project.   Figure: make sure you can connect to The upgraded projects and that you can see all of the files. Figure: Team Web Access is there and working. Note that for Team Web Access you now use the same port and URL as for TFS 2010. So in this case as I am running on the local box you need to use http://localhost:8080/tfs which will redirect you to http://localhost:8080/tfs/web for the web access. If you need to connect with a Visual Studio 2008 client you will need to use the full path of the new collection, http://[servername]/tfs/[collectionname] and this will work with all of your collections. With Visual Studio 2005 you will only be able to connect to the Default collection and in both VS2008 and VS2005 you will need to install the forward compatibility updates. Visual Studio Team System 2005 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 Visual Studio Team System 2008 Service Pack 1 Forward Compatibility Update for Team Foundation Server 2010 To make sure that you have everything up to date, make sure that you run SSW Diagnostics and get all green ticks. Upgrade Done! At this point you can send out a notice to everyone that the upgrade is complete and and give them the connection details. You need to remember that at this stage we have 2008 project upgraded to run under TFS 2010 but it is still running under that same process template that it was running before. You can only “enable” 2010 features in a process template you can’t upgrade. So what to do? Well, you need to create a new project and migrate things you want to keep across. Souse code is easy, you can move or Branch, but Work Items are more difficult as you can’t move them between projects. This instance is complicated more as the old project uses the Conchango/EMC Scrum for Team System template and I will need to write a script/application to get the work items across with their attachments in tact. That is my next task! Technorati Tags: TFS 2010,TFS 2008,VS ALM

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  • Mapping UrlEncoded POST Values in ASP.NET Web API

    - by Rick Strahl
    If there's one thing that's a bit unexpected in ASP.NET Web API, it's the limited support for mapping url encoded POST data values to simple parameters of ApiController methods. When I first looked at this I thought I was doing something wrong, because it seems mighty odd that you can bind query string values to parameters by name, but can't bind POST values to parameters in the same way. To demonstrate here's a simple example. If you have a Web API method like this:[HttpGet] public HttpResponseMessage Authenticate(string username, string password) { …} and then hit with a URL like this: http://localhost:88/samples/authenticate?Username=ricks&Password=sekrit it works just fine. The query string values are mapped to the username and password parameters of our API method. But if you now change the method to work with [HttpPost] instead like this:[HttpPost] public HttpResponseMessage Authenticate(string username, string password) { …} and hit it with a POST HTTP Request like this: POST http://localhost:88/samples/authenticate HTTP/1.1 Host: localhost:88 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Content-type: application/x-www-form-urlencoded Content-Length: 30 Username=ricks&Password=sekrit you'll find that while the request works, it doesn't actually receive the two string parameters. The username and password parameters are null and so the method is definitely going to fail. When I mentioned this over Twitter a few days ago I got a lot of responses back of why I'd want to do this in the first place - after all HTML Form submissions are the domain of MVC and not WebAPI which is a valid point. However, the more common use case is using POST Variables with AJAX calls. The following is quite common for passing simple values:$.post(url,{ Username: "Rick", Password: "sekrit" },function(result) {…}); but alas that doesn't work. How ASP.NET Web API handles Content Bodies Web API supports parsing content data in a variety of ways, but it does not deal with multiple posted content values. In effect you can only post a single content value to a Web API Action method. That one parameter can be very complex and you can bind it in a variety of ways, but ultimately you're tied to a single POST content value in your parameter definition. While it's possible to support multiple parameters on a POST/PUT operation, only one parameter can be mapped to the actual content - the rest have to be mapped to route values or the query string. Web API treats the whole request body as one big chunk of data that is sent to a Media Type Formatter that's responsible for de-serializing the content into whatever value the method requires. The restriction comes from async nature of Web API where the request data is read only once inside of the formatter that retrieves and deserializes it. Because it's read once, checking for content (like individual POST variables) first is not possible. However, Web API does provide a couple of ways to access the form POST data: Model Binding - object property mapping to bind POST values FormDataCollection - collection of POST keys/values ModelBinding POST Values - Binding POST data to Object Properties The recommended way to handle POST values in Web API is to use Model Binding, which maps individual urlencoded POST values to properties of a model object provided as the parameter. Model binding requires a single object as input to be bound to the POST data, with each POST key that matches a property name (including nested properties like Address.Street) being mapped and updated including automatic type conversion of simple types. This is a very nice feature - and a familiar one from MVC - that makes it very easy to have model objects mapped directly from inbound data. The obvious drawback with Model Binding is that you need a model for it to work: You have to provide a strongly typed object that can receive the data and this object has to map the inbound data. To rewrite the example above to use ModelBinding I have to create a class maps the properties that I need as parameters:public class LoginData { public string Username { get; set; } public string Password { get; set; } } and then accept the data like this in the API method:[HttpPost] public HttpResponseMessage Authenticate(LoginData login) { string username = login.Username; string password = login.Password; … } This works fine mapping the POST values to the properties of the login object. As a side benefit of this method definition, the method now also allows posting of JSON or XML to the same endpoint. If I change my request to send JSON like this: POST http://localhost:88/samples/authenticate HTTP/1.1 Host: localhost:88 Accept: application/jsonContent-type: application/json Content-Length: 40 {"Username":"ricks","Password":"sekrit"} it works as well and transparently, courtesy of the nice Content Negotiation features of Web API. There's nothing wrong with using Model binding and in fact it's a common practice to use (view) model object for inputs coming back from the client and mapping them into these models. But it can be  kind of a hassle if you have AJAX applications with a ton of backend hits, especially if many methods are very atomic and focused and don't effectively require a model or view. Not always do you have to pass structured data, but sometimes there are just a couple of simple response values that need to be sent back. If all you need is to pass a couple operational parameters, creating a view model object just for parameter purposes seems like overkill. Maybe you can use the query string instead (if that makes sense), but if you can't then you can often end up with a plethora of 'message objects' that serve no further  purpose than to make Model Binding work. Note that you can accept multiple parameters with ModelBinding so the following would still work:[HttpPost] public HttpResponseMessage Authenticate(LoginData login, string loginDomain) but only the object will be bound to POST data. As long as loginDomain comes from the querystring or route data this will work. Collecting POST values with FormDataCollection Another more dynamic approach to handle POST values is to collect POST data into a FormDataCollection. FormDataCollection is a very basic key/value collection (like FormCollection in MVC and Request.Form in ASP.NET in general) and then read the values out individually by querying each. [HttpPost] public HttpResponseMessage Authenticate(FormDataCollection form) { var username = form.Get("Username"); var password = form.Get("Password"); …} The downside to this approach is that it's not strongly typed, you have to handle type conversions on non-string parameters, and it gets a bit more complicated to test such as setup as you have to seed a FormDataCollection with data. On the other hand it's flexible and easy to use and especially with string parameters is easy to deal with. It's also dynamic, so if the client sends you a variety of combinations of values on which you make operating decisions, this is much easier to work with than a strongly typed object that would have to account for all possible values up front. The downside is that the code looks old school and isn't as self-documenting as a parameter list or object parameter would be. Nevertheless it's totally functionality and a viable choice for collecting POST values. What about [FromBody]? Web API also has a [FromBody] attribute that can be assigned to parameters. If you have multiple parameters on a Web API method signature you can use [FromBody] to specify which one will be parsed from the POST content. Unfortunately it's not terribly useful as it only returns content in raw format and requires a totally non-standard format ("=content") to specify your content. For more info in how FromBody works and several related issues to how POST data is mapped, you can check out Mike Stalls post: How WebAPI does Parameter Binding Not really sure where the Web API team thought [FromBody] would really be a good fit other than a down and dirty way to send a full string buffer. Extending Web API to make multiple POST Vars work? Don't think so Clearly there's no native support for multiple POST variables being mapped to parameters, which is a bit of a bummer. I know in my own work on one project my customer actually found this to be a real sticking point in their AJAX backend work, and we ended up not using Web API and using MVC JSON features instead. That's kind of sad because Web API is supposed to be the proper solution for AJAX backends. With all of ASP.NET Web API's extensibility you'd think there would be some way to build this functionality on our own, but after spending a bit of time digging and asking some of the experts from the team and Web API community I didn't hear anything that even suggests that this is possible. From what I could find I'd say it's not possible primarily because Web API's Routing engine does not account for the POST variable mapping. This means [HttpPost] methods with url encoded POST buffers are not mapped to the parameters of the endpoint, and so the routes would never even trigger a request that could be intercepted. Once the routing doesn't work there's not much that can be done. If somebody has an idea how this could be accomplished I would love to hear about it. Do we really need multi-value POST mapping? I think that that POST value mapping is a feature that one would expect of any API tool to have. If you look at common APIs out there like Flicker and Google Maps etc. they all work with POST data. POST data is very prominent much more so than JSON inputs and so supporting as many options that enable would seem to be crucial. All that aside, Web API does provide very nice features with Model Binding that allows you to capture many POST variables easily enough, and logistically this will let you build whatever you need with POST data of all shapes as long as you map objects. But having to have an object for every operation that receives a data input is going to take its toll in heavy AJAX applications, with a lot of types created that do nothing more than act as parameter containers. I also think that POST variable mapping is an expected behavior and Web APIs non-support will likely result in many, many questions like this one: How do I bind a simple POST value in ASP.NET WebAPI RC? with no clear answer to this question. I hope for V.next of WebAPI Microsoft will consider this a feature that's worth adding. Related Articles Passing multiple POST parameters to Web API Controller Methods Mike Stall's post: How Web API does Parameter Binding Where does ASP.NET Web API Fit?© Rick Strahl, West Wind Technologies, 2005-2012Posted in Web Api   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • The How-To Geek Holiday Gift Guide (Geeky Stuff We Like)

    - by The Geek
    Welcome to the very first How-To Geek Holiday Gift Guide, where we’ve put together a list of our absolute favorites to help you weed through all of the junk out there to pick the perfect gift for anybody. Though really, it’s just a list of the geeky stuff we want. We’ve got a whole range of items on the list, from cheaper gifts that most anybody can afford, to the really expensive stuff that we’re pretty sure nobody is giving us. Stocking Stuffers Here’s a couple of ideas for items that won’t break the bank. LED Keychain Micro-Light   Magcraft 1/8-Inch Rare Earth Cube Magnets Best little LED keychain light around. If they don’t need the penknife of the above item this is the perfect gift. I give them out by the handfuls and nobody ever says anything but good things about them. I’ve got ones that are years old and still running on the same battery.  Price: $8   Geeks cannot resist magnets. Jason bought this pack for his fridge because he was sick of big clunky magnets… these things are amazing. One tiny magnet, smaller than an Altoid mint, can practically hold a clipboard right to the fridge. Amazing. I spend more time playing with them on the counter than I do actually hanging stuff.  Price: $10 Lots of Geeky Mugs   Astronomy Powerful Green Laser Pointer There’s loads of fun, geeky mugs you can find on Amazon or anywhere else—and they are great choices for the geek who loves their coffee. You can get the Caffeine mug pictured here, or go with an Atari one, Canon Lens, or the Aperture mug based on Portal. Your choice. Price: $7   No, it’s not a light saber, but it’s nearly bright enough to be one—you can illuminate low flying clouds at night or just blind some aliens on your day off. All that for an extremely low price. Loads of fun. Price: $15       Geeky TV Shows and Books Sometimes you just want to relax and enjoy a some TV or a good book. Here’s a few choices. The IT Crowd Fourth Season   Doctor Who, Complete Fifth Series Ridiculous, funny show about nerds in the IT department, loved by almost all the geeks here at HTG. Justin even makes this required watching for new hires in his office so they’ll get his jokes. You can pre-order the fourth season, or pick up seasons one, two, or three for even cheaper. Price: $13   It doesn’t get any more nerdy than Eric’s pick, the fifth all-new series of Doctor Who, where the Daleks are hatching a new master plan from the heart of war-torn London. There’s also alien vampires, humanoid reptiles, and a lot more. Price: $52 Battlestar Galactica Complete Series   MAKE: Electronics: Learning Through Discovery Watch the epic fight to save the human race by finding the fabled planet Earth while being hunted by the robotic Cylons. You can grab the entire series on DVD or Blu-ray, or get the seasons individually. This isn’t your average sci-fi TV show. Price: $150 for Blu-ray.   Want to learn the fundamentals of electronics in a fun, hands-on way? The Make:Electronics book helps you build the circuits and learn how it all works—as if you had any more time between all that registry hacking and loading software on your new PC. Price: $21       Geeky Gadgets for the Gadget-Loving Geek Here’s a few of the items on our gadget list, though lets be honest: geeks are going to love almost any gadget, especially shiny new ones. Klipsch Image S4i Premium Noise-Isolating Headset with 3-Button Apple Control   GP2X Caanoo MAME/Console Emulator If you’re a real music geek looking for some serious quality in the headset for your iPhone or iPod, this is the pair that Alex recommends. They aren’t terribly cheap, but you can get the less expensive S3 earphones instead if you prefer. Price: $50-100   Eric says: “As an owner of an older version, I can say the GP2X is one of my favorite gadgets ever. Touted a “Retro Emulation Juggernaut,” GP2X runs Linux and may be the only open source software console available. Sounds too good to be true, but isn’t.” Price: $150 Roku XDS Streaming Player 1080p   Western Digital WD TV Live Plus HD Media Player If you do a lot of streaming over Netflix, Hulu Plus, Amazon’s Video on Demand, Pandora, and others, the Roku box is a great choice to get your content on your TV without paying a lot of money.  It’s also got Wireless-N built in, and it supports full 1080P HD. Price: $99   If you’ve got a home media collection sitting on a hard drive or a network server, the Western Digital box is probably the cheapest way to get that content on your TV, and it even supports Netflix streaming too. It’ll play loads of formats in full HD quality. Price: $99 Fujitsu ScanSnap S300 Color Mobile Scanner   Doxie, the amazing scanner for documents Trevor said: “This wonderful little scanner has become absolutely essential to me. My desk used to just be a gigantic pile of papers that I didn’t need at the moment, but couldn’t throw away ‘just in case.’ Now, every few weeks, I’ll run that paper pile through this and then happily shred the originals!” Price: $300   If you don’t scan quite as often and are looking for a budget scanner you can throw into your bag, or toss into a drawer in your desk, the Doxie scanner is a great alternative that I’ve been using for a while. It’s half the price, and while it’s not as full-featured as the Fujitsu, it might be a better choice for the very casual user. Price: $150       (Expensive) Gadgets Almost Anybody Will Love If you’re not sure that one of the more geeky presents is gonna work, here’s some gadgets that just about anybody is going to love, especially if they don’t have one already. Of course, some of these are a bit on the expensive side—but it’s a wish list, right? Amazon Kindle       The Kindle weighs less than a paperback book, the screen is amazing and easy on the eyes, and get ready for the kicker: the battery lasts at least a month. We aren’t kidding, either—it really lasts that long. If you don’t feel like spending money for books, you can use it to read PDFs, and if you want to get really geeky, you can hack it for custom screensavers. Price: $139 iPod Touch or iPad       You can’t go wrong with either of these presents—the iPod Touch can do almost everything the iPhone can do, including games, apps, and music, and it has the same Retina display as the iPhone, HD video recording, and a front-facing camera so you can use FaceTime. Price: $229+, depending on model. The iPad is a great tablet for playing games, browsing the web, or just using on your coffee table for guests. It’s well worth buying one—but if you’re buying for yourself, keep in mind that the iPad 2 is probably coming out in 3 months. Price: $500+ MacBook Air  The MacBook Air comes in 11” or 13” versions, and it’s an amazing little machine. It’s lightweight, the battery lasts nearly forever, and it resumes from sleep almost instantly. Since it uses an SSD drive instead of a hard drive, you’re barely going to notice any speed problems for general use. So if you’ve got a lot of money to blow, this is a killer gift. Price: $999 and up. Stuck with No Idea for a Present? Gift Cards! Yeah, you’re not going to win any “thoughtful present” awards with these, but you might just give somebody what they really want—the new Angry Birds HD for their iPad, Cut the Rope, or anything else they want. ITunes Gift Card   Amazon.com Gift Card Somebody in your circle getting a new iPod, iPhone, or iPad? You can get them an iTunes gift card, which they can use to buy music, games or apps. Yep, this way you can gift them a copy of Angry Birds if they don’t already have it. Or even Cut the Rope.   No clue what to get somebody on your list? Amazon gift cards let them buy pretty much anything they want, from organic weirdberries to big screen TVs. Yeah, it’s not as thoughtful as getting them a nice present, but look at the bright side: maybe they’ll get you an Amazon gift card and it’ll balance out. That’s the highlights from our lists—got anything else to add? Share your geeky gift ideas in the comments. Latest Features How-To Geek ETC The How-To Geek Holiday Gift Guide (Geeky Stuff We Like) LCD? LED? Plasma? The How-To Geek Guide to HDTV Technology The How-To Geek Guide to Learning Photoshop, Part 8: Filters Improve Digital Photography by Calibrating Your Monitor Our Favorite Tech: What We’re Thankful For at How-To Geek The How-To Geek Guide to Learning Photoshop, Part 7: Design and Typography Happy Snow Bears Theme for Chrome and Iron [Holiday] Download Full Command and Conquer: Tiberian Sun Game for Free Scorched Cometary Planet Wallpaper Quick Fix: Add the RSS Button Back to the Firefox Awesome Bar Dropbox Desktop Client 1.0.0 RC for Windows, Linux, and Mac Released Hang in There Scrat! – Ice Age Wallpaper

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  • SQLAuthority Book Review – DBA Survivor: Become a Rock Star DBA

    - by pinaldave
    DBA Survivor: Become a Rock Star DBA – Thomas LaRock Link to Amazon Link to Flipkart First of all, I thank all my readers when I wrote that I could not get this book in any local book stores, because they offered me to send a copy of this good book. A very special mention goes to Sripada and Jayesh for they gave so much effort in finding my home address and sending me the hard copy. Before, I did not have the copy of the book, but now I have two of it already! It surprises me how my readers were able to find my home address, which I have not publicly shared. Quick Review: This is indeed a one easy-to-read and fun book. We all work day and night with technology yet we should not forget to show our love and care for our family at home. For our souls that starve for peace and guidance, this one book is the “it” book for all the technology enthusiasts. Though this book was specifically written for DBAs, the reach is not limited to DBAs only because the lessons incorporated in it actually applies to all. This is one of the most motivating technical books I have read. Detailed Review: Let us go over a few questions first: Who wants to be as famous as rockstars in the field of Database Administration? How can one learn what it takes to become a top notch software developer? If you are a beginner in your field, how will you go to next level? Your boss may be very kind or like Dilbert’s Boss, what will you do? How do you keep growing when Eco-system around you does not support you? You are almost at top but there is someone else at the TOP, what do you do and how do you avoid office politics? As a database developer what should be your basic responsibility? and many more… I was able to completely read book in one sitting and I loved it. Before I continue with my opinion, I want to echo the opinion of Kevin Kline who has written the Forward of the book. He has truly suggested that “You hold in your hands a collection of insights and wisdom on the topic of database administration gained through many years of hard-won experience, long nights of study, and direct mentorship under some of the industry’s most talented database professionals and information technology (IT) experts.” Today, IT field is getting bigger and better, while talking about terabytes of the database becomes “more” normal every single day. The gods and demigods of database professionals are taking care of these large scale databases and are carefully maintaining them. In this world, there are only a few beginnings on the first step. There are many experts in different technology fields who are asked to address the issues with databases. There is YOU and ME, who is just new to this work. So we ask ourselves WHERE to begin and HOW to begin. We adore and follow the religion of our rockstars, but oftentimes we really have no idea about their background and their struggles. Every rockstar has his success story which needs to be digested before learning his tricks and tips. This book starts with the same note and teaches the two most important lessons for anybody who wants to be a DBA Rockstar –  to focus on their single goal of learning and to excel the technology. The story starts with three simple guidelines – Get Prepared, Get Trained, Get Certified. Once a person learns the skills, and then, it would be about time that he needs to enrich or to improve those skills you have learned. I am sure that the right opportunity will come finding themselves and they will not have to go run behind it. However, the real challenge for any person is the first day or first week. A new employee, no matter how much experienced he is, sometimes has no clue about what should one do at new job. Chapter 2 and chapter 3 precisely talk about what one should do as soon as the new job begins. It is also written with keeping the fact in focus that each job can be very much different but there are few infrastructure setups and programming concepts are the same. Learning basics of database was really interesting. I like to focus on the roots of any technology. It is important to understand the structure of the database before suggesting what indexes needs to be created, the same way this book covers the most essential knowledge one must learn by most database developers. I think the title of the fourth chapter is my favorite sentence in this book. I can see that I will be saying this again and again in the future – “A Development Server Is a Production Server to a Developer“. I have worked in the software industry for almost 8 years now and I have seen so many developers sitting on their chairs and waiting for instructions from their lead about how to improve the code or what to do the next. When I talk to them, I suggest that the experiment with their server and try various techniques. I think they all should understand that for them, a development server is their production server and needs to pay proper attention to the code from the beginning. There should be NO any inappropriate code from the beginning. One has to fully focus and give their best, if they are not sure they should ask but should do something and stay active. Chapter 5 and 6 talks about two essential skills for any developer and database administration – what are the ethics of developers when they are working with production server and how to support software which is running on the production server. I have met many people who know the theory by heart but when put in front of keyboard they do not know where to start. The first thing they do opening the browser and searching online, instead of opening SQL Server Management Studio. This can very well happen to anybody who is experienced as well. Chapter 5 and 6 addresses that situation as well includes the handy scripts which can solve almost all the basic trouble shooting issues. “Where’s the Buffet?” By far, this is the best chapter in this book. If you have ever met me, you would know that I love food. I think after reading this chapter, I felt Thomas has written this just keeping me in mind. I think there will be many other people who feel the same way, too. Even my wife who read this chapter thought this was specifically written for me. I will not talk any more about this chapter as this is one must read chapter. And of course this is about real ‘FOOD‘. I am an SQL Server Trainer and Consultant and I totally agree with the point made in the chapter 8 of this book. Yes, it says here that what is necessary to train employees and people. Millions of dollars worth the labor is continuously done in the world which has faults and incorrect. Once something goes wrong, very expensive consultant comes in and fixes the problem. This whole cycle which can be stopped and improved if proper training is done. There is plenty of free trainings available as well, if one cannot afford paid training. “Connect. Learn. Share” – I think this is a great summary and bird’s eye view of this book. Networking is the key. Everything which is discussed in this book can be taken to next level if one properly uses this tips and continuously grow with it. Connecting with others, helping learn each other and building the good knowledge sharing environment should be the goal of everyone. Before I end the review I want to share a real experience. I have personally met one DBA who has worked in a single department in a company for so long that when he was put in a different department in his company due to closing that department, he could not adjust and quit the job despite the same people and company around him. Adjusting in the new environment gets much tougher as one person gets more and more experienced. This book precisely addresses the same issue along with their solutions. I just cannot stop comparing the book with my personal journey. I found so many things which are coincidently in the book is written as how we developer and DBA think. I must express special thanks to Thomas for taking time in his personal life and write this book for us. This book is indeed a book for everybody who wants to grow healthy in the tough and competitive environment. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, SQLAuthority Book Review, SQLAuthority News, SQLServer, T SQL, Technology

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  • Back to Basics: When does a .NET Assembly Dependency get loaded

    - by Rick Strahl
    When we work on typical day to day applications, it's easy to forget some of the core features of the .NET framework. For me personally it's been a long time since I've learned about some of the underlying CLR system level services even though I rely on them on a daily basis. I often think only about high level application constructs and/or high level framework functionality, but the low level stuff is often just taken for granted. Over the last week at DevConnections I had all sorts of low level discussions with other developers about the inner workings of this or that technology (especially in light of my Low Level ASP.NET Architecture talk and the Razor Hosting talk). One topic that came up a couple of times and ended up a point of confusion even amongst some seasoned developers (including some folks from Microsoft <snicker>) is when assemblies actually load into a .NET process. There are a number of different ways that assemblies are loaded in .NET. When you create a typical project assemblies usually come from: The Assembly reference list of the top level 'executable' project The Assembly references of referenced projects Dynamically loaded at runtime via AppDomain/Reflection loading In addition .NET automatically loads mscorlib (most of the System namespace) the boot process that hosts the .NET runtime in EXE apps, or some other kind of runtime hosting environment (runtime hosting in servers like IIS, SQL Server or COM Interop). In hosting environments the runtime host may also pre-load a bunch of assemblies on its own (for example the ASP.NET host requires all sorts of assemblies just to run itself, before ever routing into your user specific code). Assembly Loading The most obvious source of loaded assemblies is the top level application's assembly reference list. You can add assembly references to a top level application and those assembly references are then available to the application. In a nutshell, referenced assemblies are not immediately loaded - they are loaded on the fly as needed. So regardless of whether you have an assembly reference in a top level project, or a dependent assembly assemblies typically load on an as needed basis, unless explicitly loaded by user code. The same is true of dependent assemblies. To check this out I ran a simple test: I have a utility assembly Westwind.Utilities which is a general purpose library that can work in any type of project. Due to a couple of small requirements for encoding and a logging piece that allows logging Web content (dependency on HttpContext.Current) this utility library has a dependency on System.Web. Now System.Web is a pretty large assembly and generally you'd want to avoid adding it to a non-Web project if it can be helped. So I created a Console Application that loads my utility library: You can see that the top level Console app a reference to Westwind.Utilities and System.Data (beyond the core .NET libs). The Westwind.Utilities project on the other hand has quite a few dependencies including System.Web. I then add a main program that accesses only a simple utillity method in the Westwind.Utilities library that doesn't require any of the classes that access System.Web: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.ReadLine(); } StringUtils.NewStringId() calls into Westwind.Utilities, but it doesn't rely on System.Web. Any guesses what the assembly list looks like when I stop the code on the ReadLine() command? I'll wait here while you think about it… … … So, when I stop on ReadLine() and then fire up Process Explorer and check the assembly list I get: We can see here that .NET has not actually loaded any of the dependencies of the Westwind.Utilities assembly. Also not loaded is the top level System.Data reference even though it's in the dependent assembly list of the top level project. Since this particular function I called only uses core System functionality (contained in mscorlib) there's in fact nothing else loaded beyond the main application and my Westwind.Utilities assembly that contains the method accessed. None of the dependencies of Westwind.Utilities loaded. If you were to open the assembly in a disassembler like Reflector or ILSpy, you would however see all the compiled in dependencies. The referenced assemblies are in the dependency list and they are loadable, but they are not immediately loaded by the application. In other words the C# compiler and .NET linker are smart enough to figure out the dependencies based on the code that actually is referenced from your application and any dependencies cascading down into the dependencies from your top level application into the referenced assemblies. In the example above the usage requirement is pretty obvious since I'm only calling a single static method and then exiting the app, but in more complex applications these dependency relationships become very complicated - however it's all taken care of by the compiler and linker figuring out what types and members are actually referenced and including only those assemblies that are in fact referenced in your code or required by any of your dependencies. The good news here is: That if you are referencing an assembly that has a dependency on something like System.Web in a few places that are not actually accessed by any of your code or any dependent assembly code that you are calling, that assembly is never loaded into memory! Some Hosting Environments pre-load Assemblies The load behavior can vary however. In Console and desktop applications we have full control over assembly loading so we see the core CLR behavior. However other environments like ASP.NET for example will preload referenced assemblies explicitly as part of the startup process - primarily to minimize load conflicts. Specifically ASP.NET pre-loads all assemblies referenced in the assembly list and the /bin folder. So in Web applications it definitely pays to minimize your top level assemblies if they are not used. Understanding when Assemblies Load To clarify and see it actually happen what I described in the first example , let's look at a couple of other scenarios. To see assemblies loading at runtime in real time lets create a utility function to print out loaded assemblies to the console: public static void PrintAssemblies() { var assemblies = AppDomain.CurrentDomain.GetAssemblies(); foreach (var assembly in assemblies) { Console.WriteLine(assembly.GetName()); } } Now let's look at the first scenario where I have class method that references internally uses System.Web. In the first scenario lets add a method to my main program like this: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.ReadLine(); PrintAssemblies(); } public static void WebLogEntry() { var entry = new WebLogEntry(); entry.UpdateFromRequest(); Console.WriteLine(entry.QueryString); } UpdateFromWebRequest() internally accesses HttpContext.Current to read some information of the ASP.NET Request object so it clearly needs a reference System.Web to work. In this first example, the method that holds the calling code is never called, but exists as a static method that can potentially be called externally at some point. What do you think will happen here with the assembly loading? Will System.Web load in this example? No - it doesn't. Because the WebLogEntry() method is never called by the mainline application (or anywhere else) System.Web is not loaded. .NET dynamically loads assemblies as code that needs it is called. No code references the WebLogEntry() method and so System.Web is never loaded. Next, let's add the call to this method, which should trigger System.Web to be loaded because a dependency exists. Let's change the code to: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.WriteLine("--- Before:"); PrintAssemblies(); WebLogEntry(); Console.WriteLine("--- After:"); PrintAssemblies(); Console.ReadLine(); } public static void WebLogEntry() { var entry = new WebLogEntry(); entry.UpdateFromRequest(); Console.WriteLine(entry.QueryString); } Looking at the code now, when do you think System.Web will be loaded? Will the before list include it? Yup System.Web gets loaded, but only after it's actually referenced. In fact, just until before the call to UpdateFromRequest() System.Web is not loaded - it only loads when the method is actually called and requires the reference in the executing code. Moral of the Story So what have we learned - or maybe remembered again? Dependent Assembly References are not pre-loaded when an application starts (by default) Dependent Assemblies that are not referenced by executing code are never loaded Dependent Assemblies are just in time loaded when first referenced in code All of this is nothing new - .NET has always worked like this. But it's good to have a refresher now and then and go through the exercise of seeing it work in action. It's not one of those things we think about everyday, and as I found out last week, I couldn't remember exactly how it worked since it's been so long since I've learned about this. And apparently I'm not the only one as several other people I had discussions with in relation to loaded assemblies also didn't recall exactly what should happen or assumed incorrectly that just having a reference automatically loads the assembly. The moral of the story for me is: Trying at all costs to eliminate an assembly reference from a component is not quite as important as it's often made out to be. For example, the Westwind.Utilities module described above has a logging component, including a Web specific logging entry that supports pulling information from the active HTTP Context. Adding that feature requires a reference to System.Web. Should I worry about this in the scope of this library? Probably not, because if I don't use that one class of nearly a hundred, System.Web never gets pulled into the parent process. IOW, System.Web only loads when I use that specific feature and if I am, well I clearly have to be running in a Web environment anyway to use it realistically. The alternative would be considerably uglier: Pulling out the WebLogEntry class and sticking it into another assembly and breaking up the logging code. In this case - definitely not worth it. So, .NET definitely goes through some pretty nifty optimizations to ensure that it loads only what it needs and in most cases you can just rely on .NET to do the right thing. Sometimes though assembly loading can go wrong (especially when signed and versioned local assemblies are involved), but that's subject for a whole other post…© Rick Strahl, West Wind Technologies, 2005-2012Posted in .NET  CSharp   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • How to Reuse Your Old Wi-Fi Router as a Network Switch

    - by Jason Fitzpatrick
    Just because your old Wi-Fi router has been replaced by a newer model doesn’t mean it needs to gather dust in the closet. Read on as we show you how to take an old and underpowered Wi-Fi router and turn it into a respectable network switch (saving your $20 in the process). Image by mmgallan. Why Do I Want To Do This? Wi-Fi technology has changed significantly in the last ten years but Ethernet-based networking has changed very little. As such, a Wi-Fi router with 2006-era guts is lagging significantly behind current Wi-Fi router technology, but the Ethernet networking component of the device is just as useful as ever; aside from potentially being only 100Mbs instead of 1000Mbs capable (which for 99% of home applications is irrelevant) Ethernet is Ethernet. What does this matter to you, the consumer? It means that even though your old router doesn’t hack it for your Wi-Fi needs any longer the device is still a perfectly serviceable (and high quality) network switch. When do you need a network switch? Any time you want to share an Ethernet cable among multiple devices, you need a switch. For example, let’s say you have a single Ethernet wall jack behind your entertainment center. Unfortunately you have four devices that you want to link to your local network via hardline including your smart HDTV, DVR, Xbox, and a little Raspberry Pi running XBMC. Instead of spending $20-30 to purchase a brand new switch of comparable build quality to your old Wi-Fi router it makes financial sense (and is environmentally friendly) to invest five minutes of your time tweaking the settings on the old router to turn it from a Wi-Fi access point and routing tool into a network switch–perfect for dropping behind your entertainment center so that your DVR, Xbox, and media center computer can all share an Ethernet connection. What Do I Need? For this tutorial you’ll need a few things, all of which you likely have readily on hand or are free for download. To follow the basic portion of the tutorial, you’ll need the following: 1 Wi-Fi router with Ethernet ports 1 Computer with Ethernet jack 1 Ethernet cable For the advanced tutorial you’ll need all of those things, plus: 1 copy of DD-WRT firmware for your Wi-Fi router We’re conducting the experiment with a Linksys WRT54GL Wi-Fi router. The WRT54 series is one of the best selling Wi-Fi router series of all time and there’s a good chance a significant number of readers have one (or more) of them stuffed in an office closet. Even if you don’t have one of the WRT54 series routers, however, the principles we’re outlining here apply to all Wi-Fi routers; as long as your router administration panel allows the necessary changes you can follow right along with us. A quick note on the difference between the basic and advanced versions of this tutorial before we proceed. Your typical Wi-Fi router has 5 Ethernet ports on the back: 1 labeled “Internet”, “WAN”, or a variation thereof and intended to be connected to your DSL/Cable modem, and 4 labeled 1-4 intended to connect Ethernet devices like computers, printers, and game consoles directly to the Wi-Fi router. When you convert a Wi-Fi router to a switch, in most situations, you’ll lose two port as the “Internet” port cannot be used as a normal switch port and one of the switch ports becomes the input port for the Ethernet cable linking the switch to the main network. This means, referencing the diagram above, you’d lose the WAN port and LAN port 1, but retain LAN ports 2, 3, and 4 for use. If you only need to switch for 2-3 devices this may be satisfactory. However, for those of you that would prefer a more traditional switch setup where there is a dedicated WAN port and the rest of the ports are accessible, you’ll need to flash a third-party router firmware like the powerful DD-WRT onto your device. Doing so opens up the router to a greater degree of modification and allows you to assign the previously reserved WAN port to the switch, thus opening up LAN ports 1-4. Even if you don’t intend to use that extra port, DD-WRT offers you so many more options that it’s worth the extra few steps. Preparing Your Router for Life as a Switch Before we jump right in to shutting down the Wi-Fi functionality and repurposing your device as a network switch, there are a few important prep steps to attend to. First, you want to reset the router (if you just flashed a new firmware to your router, skip this step). Following the reset procedures for your particular router or go with what is known as the “Peacock Method” wherein you hold down the reset button for thirty seconds, unplug the router and wait (while still holding the reset button) for thirty seconds, and then plug it in while, again, continuing to hold down the rest button. Over the life of a router there are a variety of changes made, big and small, so it’s best to wipe them all back to the factory default before repurposing the router as a switch. Second, after resetting, we need to change the IP address of the device on the local network to an address which does not directly conflict with the new router. The typical default IP address for a home router is 192.168.1.1; if you ever need to get back into the administration panel of the router-turned-switch to check on things or make changes it will be a real hassle if the IP address of the device conflicts with the new home router. The simplest way to deal with this is to assign an address close to the actual router address but outside the range of addresses that your router will assign via the DHCP client; a good pick then is 192.168.1.2. Once the router is reset (or re-flashed) and has been assigned a new IP address, it’s time to configure it as a switch. Basic Router to Switch Configuration If you don’t want to (or need to) flash new firmware onto your device to open up that extra port, this is the section of the tutorial for you: we’ll cover how to take a stock router, our previously mentioned WRT54 series Linksys, and convert it to a switch. Hook the Wi-Fi router up to the network via one of the LAN ports (consider the WAN port as good as dead from this point forward, unless you start using the router in its traditional function again or later flash a more advanced firmware to the device, the port is officially retired at this point). Open the administration control panel via  web browser on a connected computer. Before we get started two things: first,  anything we don’t explicitly instruct you to change should be left in the default factory-reset setting as you find it, and two, change the settings in the order we list them as some settings can’t be changed after certain features are disabled. To start, let’s navigate to Setup ->Basic Setup. Here you need to change the following things: Local IP Address: [different than the primary router, e.g. 192.168.1.2] Subnet Mask: [same as the primary router, e.g. 255.255.255.0] DHCP Server: Disable Save with the “Save Settings” button and then navigate to Setup -> Advanced Routing: Operating Mode: Router This particular setting is very counterintuitive. The “Operating Mode” toggle tells the device whether or not it should enable the Network Address Translation (NAT)  feature. Because we’re turning a smart piece of networking hardware into a relatively dumb one, we don’t need this feature so we switch from Gateway mode (NAT on) to Router mode (NAT off). Our next stop is Wireless -> Basic Wireless Settings: Wireless SSID Broadcast: Disable Wireless Network Mode: Disabled After disabling the wireless we’re going to, again, do something counterintuitive. Navigate to Wireless -> Wireless Security and set the following parameters: Security Mode: WPA2 Personal WPA Algorithms: TKIP+AES WPA Shared Key: [select some random string of letters, numbers, and symbols like JF#d$di!Hdgio890] Now you may be asking yourself, why on Earth are we setting a rather secure Wi-Fi configuration on a Wi-Fi router we’re not going to use as a Wi-Fi node? On the off chance that something strange happens after, say, a power outage when your router-turned-switch cycles on and off a bunch of times and the Wi-Fi functionality is activated we don’t want to be running the Wi-Fi node wide open and granting unfettered access to your network. While the chances of this are next-to-nonexistent, it takes only a few seconds to apply the security measure so there’s little reason not to. Save your changes and navigate to Security ->Firewall. Uncheck everything but Filter Multicast Firewall Protect: Disable At this point you can save your changes again, review the changes you’ve made to ensure they all stuck, and then deploy your “new” switch wherever it is needed. Advanced Router to Switch Configuration For the advanced configuration, you’ll need a copy of DD-WRT installed on your router. Although doing so is an extra few steps, it gives you a lot more control over the process and liberates an extra port on the device. Hook the Wi-Fi router up to the network via one of the LAN ports (later you can switch the cable to the WAN port). Open the administration control panel via web browser on the connected computer. Navigate to the Setup -> Basic Setup tab to get started. In the Basic Setup tab, ensure the following settings are adjusted. The setting changes are not optional and are required to turn the Wi-Fi router into a switch. WAN Connection Type: Disabled Local IP Address: [different than the primary router, e.g. 192.168.1.2] Subnet Mask: [same as the primary router, e.g. 255.255.255.0] DHCP Server: Disable In addition to disabling the DHCP server, also uncheck all the DNSMasq boxes as the bottom of the DHCP sub-menu. If you want to activate the extra port (and why wouldn’t you), in the WAN port section: Assign WAN Port to Switch [X] At this point the router has become a switch and you have access to the WAN port so the LAN ports are all free. Since we’re already in the control panel, however, we might as well flip a few optional toggles that further lock down the switch and prevent something odd from happening. The optional settings are arranged via the menu you find them in. Remember to save your settings with the save button before moving onto a new tab. While still in the Setup -> Basic Setup menu, change the following: Gateway/Local DNS : [IP address of primary router, e.g. 192.168.1.1] NTP Client : Disable The next step is to turn off the radio completely (which not only kills the Wi-Fi but actually powers the physical radio chip off). Navigate to Wireless -> Advanced Settings -> Radio Time Restrictions: Radio Scheduling: Enable Select “Always Off” There’s no need to create a potential security problem by leaving the Wi-Fi radio on, the above toggle turns it completely off. Under Services -> Services: DNSMasq : Disable ttraff Daemon : Disable Under the Security -> Firewall tab, uncheck every box except “Filter Multicast”, as seen in the screenshot above, and then disable SPI Firewall. Once you’re done here save and move on to the Administration tab. Under Administration -> Management:  Info Site Password Protection : Enable Info Site MAC Masking : Disable CRON : Disable 802.1x : Disable Routing : Disable After this final round of tweaks, save and then apply your settings. Your router has now been, strategically, dumbed down enough to plod along as a very dependable little switch. Time to stuff it behind your desk or entertainment center and streamline your cabling.     

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  • TOTD #166: Using NoSQL database in your Java EE 6 Applications on GlassFish - MongoDB for now!

    - by arungupta
    The Java EE 6 platform includes Java Persistence API to work with RDBMS. The JPA specification defines a comprehensive API that includes, but not restricted to, how a database table can be mapped to a POJO and vice versa, provides mechanisms how a PersistenceContext can be injected in a @Stateless bean and then be used for performing different operations on the database table and write typesafe queries. There are several well known advantages of RDBMS but the NoSQL movement has gained traction over past couple of years. The NoSQL databases are not intended to be a replacement for the mainstream RDBMS. As Philosophy of NoSQL explains, NoSQL database was designed for casual use where all the features typically provided by an RDBMS are not required. The name "NoSQL" is more of a category of databases that is more known for what it is not rather than what it is. The basic principles of NoSQL database are: No need to have a pre-defined schema and that makes them a schema-less database. Addition of new properties to existing objects is easy and does not require ALTER TABLE. The unstructured data gives flexibility to change the format of data any time without downtime or reduced service levels. Also there are no joins happening on the server because there is no structure and thus no relation between them. Scalability and performance is more important than the entire set of functionality typically provided by an RDBMS. This set of databases provide eventual consistency and/or transactions restricted to single items but more focus on CRUD. Not be restricted to SQL to access the information stored in the backing database. Designed to scale-out (horizontal) instead of scale-up (vertical). This is important knowing that databases, and everything else as well, is moving into the cloud. RBDMS can scale-out using sharding but requires complex management and not for the faint of heart. Unlike RBDMS which require a separate caching tier, most of the NoSQL databases comes with integrated caching. Designed for less management and simpler data models lead to lower administration as well. There are primarily three types of NoSQL databases: Key-Value stores (e.g. Cassandra and Riak) Document databases (MongoDB or CouchDB) Graph databases (Neo4J) You may think NoSQL is panacea but as I mentioned above they are not meant to replace the mainstream databases and here is why: RDBMS have been around for many years, very stable, and functionally rich. This is something CIOs and CTOs can bet their money on without much worry. There is a reason 98% of Fortune 100 companies run Oracle :-) NoSQL is cutting edge, brings excitement to developers, but enterprises are cautious about them. Commercial databases like Oracle are well supported by the backing enterprises in terms of providing support resources on a global scale. There is a full ecosystem built around these commercial databases providing training, performance tuning, architecture guidance, and everything else. NoSQL is fairly new and typically backed by a single company not able to meet the scale of these big enterprises. NoSQL databases are good for CRUDing operations but business intelligence is extremely important for enterprises to stay competitive. RDBMS provide extensive tooling to generate this data but that was not the original intention of NoSQL databases and is lacking in that area. Generating any meaningful information other than CRUDing require extensive programming. Not suited for complex transactions such as banking systems or other highly transactional applications requiring 2-phase commit. SQL cannot be used with NoSQL databases and writing simple queries can be involving. Enough talking, lets take a look at some code. This blog has published multiple blogs on how to access a RDBMS using JPA in a Java EE 6 application. This Tip Of The Day (TOTD) will show you can use MongoDB (a document-oriented database) with a typical 3-tier Java EE 6 application. Lets get started! The complete source code of this project can be downloaded here. Download MongoDB for your platform from here (1.8.2 as of this writing) and start the server as: arun@ArunUbuntu:~/tools/mongodb-linux-x86_64-1.8.2/bin$./mongod./mongod --help for help and startup optionsSun Jun 26 20:41:11 [initandlisten] MongoDB starting : pid=11210port=27017 dbpath=/data/db/ 64-bit Sun Jun 26 20:41:11 [initandlisten] db version v1.8.2, pdfile version4.5Sun Jun 26 20:41:11 [initandlisten] git version:433bbaa14aaba6860da15bd4de8edf600f56501bSun Jun 26 20:41:11 [initandlisten] build sys info: Linuxbs-linux64.10gen.cc 2.6.21.7-2.ec2.v1.2.fc8xen #1 SMP Fri Nov 2017:48:28 EST 2009 x86_64 BOOST_LIB_VERSION=1_41Sun Jun 26 20:41:11 [initandlisten] waiting for connections on port 27017Sun Jun 26 20:41:11 [websvr] web admin interface listening on port 28017 The default directory for the database is /data/db and needs to be created as: sudo mkdir -p /data/db/sudo chown `id -u` /data/db You can specify a different directory using "--dbpath" option. Refer to Quickstart for your specific platform. Using NetBeans, create a Java EE 6 project and make sure to enable CDI and add JavaServer Faces framework. Download MongoDB Java Driver (2.6.3 of this writing) and add it to the project library by selecting "Properties", "LIbraries", "Add Library...", creating a new library by specifying the location of the JAR file, and adding the library to the created project. Edit the generated "index.xhtml" such that it looks like: <h1>Add a new movie</h1><h:form> Name: <h:inputText value="#{movie.name}" size="20"/><br/> Year: <h:inputText value="#{movie.year}" size="6"/><br/> Language: <h:inputText value="#{movie.language}" size="20"/><br/> <h:commandButton actionListener="#{movieSessionBean.createMovie}" action="show" title="Add" value="submit"/></h:form> This page has a simple HTML form with three text boxes and a submit button. The text boxes take name, year, and language of a movie and the submit button invokes the "createMovie" method of "movieSessionBean" and then render "show.xhtml". Create "show.xhtml" ("New" -> "Other..." -> "Other" -> "XHTML File") such that it looks like: <head> <title><h1>List of movies</h1></title> </head> <body> <h:form> <h:dataTable value="#{movieSessionBean.movies}" var="m" > <h:column><f:facet name="header">Name</f:facet>#{m.name}</h:column> <h:column><f:facet name="header">Year</f:facet>#{m.year}</h:column> <h:column><f:facet name="header">Language</f:facet>#{m.language}</h:column> </h:dataTable> </h:form> This page shows the name, year, and language of all movies stored in the database so far. The list of movies is returned by "movieSessionBean.movies" property. Now create the "Movie" class such that it looks like: import com.mongodb.BasicDBObject;import com.mongodb.BasicDBObject;import com.mongodb.DBObject;import javax.enterprise.inject.Model;import javax.validation.constraints.Size;/** * @author arun */@Modelpublic class Movie { @Size(min=1, max=20) private String name; @Size(min=1, max=20) private String language; private int year; // getters and setters for "name", "year", "language" public BasicDBObject toDBObject() { BasicDBObject doc = new BasicDBObject(); doc.put("name", name); doc.put("year", year); doc.put("language", language); return doc; } public static Movie fromDBObject(DBObject doc) { Movie m = new Movie(); m.name = (String)doc.get("name"); m.year = (int)doc.get("year"); m.language = (String)doc.get("language"); return m; } @Override public String toString() { return name + ", " + year + ", " + language; }} Other than the usual boilerplate code, the key methods here are "toDBObject" and "fromDBObject". These methods provide a conversion from "Movie" -> "DBObject" and vice versa. The "DBObject" is a MongoDB class that comes as part of the mongo-2.6.3.jar file and which we added to our project earlier.  The complete javadoc for 2.6.3 can be seen here. Notice, this class also uses Bean Validation constraints and will be honored by the JSF layer. Finally, create "MovieSessionBean" stateless EJB with all the business logic such that it looks like: package org.glassfish.samples;import com.mongodb.BasicDBObject;import com.mongodb.DB;import com.mongodb.DBCollection;import com.mongodb.DBCursor;import com.mongodb.DBObject;import com.mongodb.Mongo;import java.net.UnknownHostException;import java.util.ArrayList;import java.util.List;import javax.annotation.PostConstruct;import javax.ejb.Stateless;import javax.inject.Inject;import javax.inject.Named;/** * @author arun */@Stateless@Namedpublic class MovieSessionBean { @Inject Movie movie; DBCollection movieColl; @PostConstruct private void initDB() throws UnknownHostException { Mongo m = new Mongo(); DB db = m.getDB("movieDB"); movieColl = db.getCollection("movies"); if (movieColl == null) { movieColl = db.createCollection("movies", null); } } public void createMovie() { BasicDBObject doc = movie.toDBObject(); movieColl.insert(doc); } public List<Movie> getMovies() { List<Movie> movies = new ArrayList(); DBCursor cur = movieColl.find(); System.out.println("getMovies: Found " + cur.size() + " movie(s)"); for (DBObject dbo : cur.toArray()) { movies.add(Movie.fromDBObject(dbo)); } return movies; }} The database is initialized in @PostConstruct. Instead of a working with a database table, NoSQL databases work with a schema-less document. The "Movie" class is the document in our case and stored in the collection "movies". The collection allows us to perform query functions on all movies. The "getMovies" method invokes "find" method on the collection which is equivalent to the SQL query "select * from movies" and then returns a List<Movie>. Also notice that there is no "persistence.xml" in the project. Right-click and run the project to see the output as: Enter some values in the text box and click on enter to see the result as: If you reached here then you've successfully used MongoDB in your Java EE 6 application, congratulations! Some food for thought and further play ... SQL to MongoDB mapping shows mapping between traditional SQL -> Mongo query language. Tutorial shows fun things you can do with MongoDB. Try the interactive online shell  The cookbook provides common ways of using MongoDB In terms of this project, here are some tasks that can be tried: Encapsulate database management in a JPA persistence provider. Is it even worth it because the capabilities are going to be very different ? MongoDB uses "BSonObject" class for JSON representation, add @XmlRootElement on a POJO and how a compatible JSON representation can be generated. This will make the fromXXX and toXXX methods redundant.

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  • TOTD #166: Using NoSQL database in your Java EE 6 Applications on GlassFish - MongoDB for now!

    - by arungupta
    The Java EE 6 platform includes Java Persistence API to work with RDBMS. The JPA specification defines a comprehensive API that includes, but not restricted to, how a database table can be mapped to a POJO and vice versa, provides mechanisms how a PersistenceContext can be injected in a @Stateless bean and then be used for performing different operations on the database table and write typesafe queries. There are several well known advantages of RDBMS but the NoSQL movement has gained traction over past couple of years. The NoSQL databases are not intended to be a replacement for the mainstream RDBMS. As Philosophy of NoSQL explains, NoSQL database was designed for casual use where all the features typically provided by an RDBMS are not required. The name "NoSQL" is more of a category of databases that is more known for what it is not rather than what it is. The basic principles of NoSQL database are: No need to have a pre-defined schema and that makes them a schema-less database. Addition of new properties to existing objects is easy and does not require ALTER TABLE. The unstructured data gives flexibility to change the format of data any time without downtime or reduced service levels. Also there are no joins happening on the server because there is no structure and thus no relation between them. Scalability and performance is more important than the entire set of functionality typically provided by an RDBMS. This set of databases provide eventual consistency and/or transactions restricted to single items but more focus on CRUD. Not be restricted to SQL to access the information stored in the backing database. Designed to scale-out (horizontal) instead of scale-up (vertical). This is important knowing that databases, and everything else as well, is moving into the cloud. RBDMS can scale-out using sharding but requires complex management and not for the faint of heart. Unlike RBDMS which require a separate caching tier, most of the NoSQL databases comes with integrated caching. Designed for less management and simpler data models lead to lower administration as well. There are primarily three types of NoSQL databases: Key-Value stores (e.g. Cassandra and Riak) Document databases (MongoDB or CouchDB) Graph databases (Neo4J) You may think NoSQL is panacea but as I mentioned above they are not meant to replace the mainstream databases and here is why: RDBMS have been around for many years, very stable, and functionally rich. This is something CIOs and CTOs can bet their money on without much worry. There is a reason 98% of Fortune 100 companies run Oracle :-) NoSQL is cutting edge, brings excitement to developers, but enterprises are cautious about them. Commercial databases like Oracle are well supported by the backing enterprises in terms of providing support resources on a global scale. There is a full ecosystem built around these commercial databases providing training, performance tuning, architecture guidance, and everything else. NoSQL is fairly new and typically backed by a single company not able to meet the scale of these big enterprises. NoSQL databases are good for CRUDing operations but business intelligence is extremely important for enterprises to stay competitive. RDBMS provide extensive tooling to generate this data but that was not the original intention of NoSQL databases and is lacking in that area. Generating any meaningful information other than CRUDing require extensive programming. Not suited for complex transactions such as banking systems or other highly transactional applications requiring 2-phase commit. SQL cannot be used with NoSQL databases and writing simple queries can be involving. Enough talking, lets take a look at some code. This blog has published multiple blogs on how to access a RDBMS using JPA in a Java EE 6 application. This Tip Of The Day (TOTD) will show you can use MongoDB (a document-oriented database) with a typical 3-tier Java EE 6 application. Lets get started! The complete source code of this project can be downloaded here. Download MongoDB for your platform from here (1.8.2 as of this writing) and start the server as: arun@ArunUbuntu:~/tools/mongodb-linux-x86_64-1.8.2/bin$./mongod./mongod --help for help and startup optionsSun Jun 26 20:41:11 [initandlisten] MongoDB starting : pid=11210port=27017 dbpath=/data/db/ 64-bit Sun Jun 26 20:41:11 [initandlisten] db version v1.8.2, pdfile version4.5Sun Jun 26 20:41:11 [initandlisten] git version:433bbaa14aaba6860da15bd4de8edf600f56501bSun Jun 26 20:41:11 [initandlisten] build sys info: Linuxbs-linux64.10gen.cc 2.6.21.7-2.ec2.v1.2.fc8xen #1 SMP Fri Nov 2017:48:28 EST 2009 x86_64 BOOST_LIB_VERSION=1_41Sun Jun 26 20:41:11 [initandlisten] waiting for connections on port 27017Sun Jun 26 20:41:11 [websvr] web admin interface listening on port 28017 The default directory for the database is /data/db and needs to be created as: sudo mkdir -p /data/db/sudo chown `id -u` /data/db You can specify a different directory using "--dbpath" option. Refer to Quickstart for your specific platform. Using NetBeans, create a Java EE 6 project and make sure to enable CDI and add JavaServer Faces framework. Download MongoDB Java Driver (2.6.3 of this writing) and add it to the project library by selecting "Properties", "LIbraries", "Add Library...", creating a new library by specifying the location of the JAR file, and adding the library to the created project. Edit the generated "index.xhtml" such that it looks like: <h1>Add a new movie</h1><h:form> Name: <h:inputText value="#{movie.name}" size="20"/><br/> Year: <h:inputText value="#{movie.year}" size="6"/><br/> Language: <h:inputText value="#{movie.language}" size="20"/><br/> <h:commandButton actionListener="#{movieSessionBean.createMovie}" action="show" title="Add" value="submit"/></h:form> This page has a simple HTML form with three text boxes and a submit button. The text boxes take name, year, and language of a movie and the submit button invokes the "createMovie" method of "movieSessionBean" and then render "show.xhtml". Create "show.xhtml" ("New" -> "Other..." -> "Other" -> "XHTML File") such that it looks like: <head> <title><h1>List of movies</h1></title> </head> <body> <h:form> <h:dataTable value="#{movieSessionBean.movies}" var="m" > <h:column><f:facet name="header">Name</f:facet>#{m.name}</h:column> <h:column><f:facet name="header">Year</f:facet>#{m.year}</h:column> <h:column><f:facet name="header">Language</f:facet>#{m.language}</h:column> </h:dataTable> </h:form> This page shows the name, year, and language of all movies stored in the database so far. The list of movies is returned by "movieSessionBean.movies" property. Now create the "Movie" class such that it looks like: import com.mongodb.BasicDBObject;import com.mongodb.BasicDBObject;import com.mongodb.DBObject;import javax.enterprise.inject.Model;import javax.validation.constraints.Size;/** * @author arun */@Modelpublic class Movie { @Size(min=1, max=20) private String name; @Size(min=1, max=20) private String language; private int year; // getters and setters for "name", "year", "language" public BasicDBObject toDBObject() { BasicDBObject doc = new BasicDBObject(); doc.put("name", name); doc.put("year", year); doc.put("language", language); return doc; } public static Movie fromDBObject(DBObject doc) { Movie m = new Movie(); m.name = (String)doc.get("name"); m.year = (int)doc.get("year"); m.language = (String)doc.get("language"); return m; } @Override public String toString() { return name + ", " + year + ", " + language; }} Other than the usual boilerplate code, the key methods here are "toDBObject" and "fromDBObject". These methods provide a conversion from "Movie" -> "DBObject" and vice versa. The "DBObject" is a MongoDB class that comes as part of the mongo-2.6.3.jar file and which we added to our project earlier.  The complete javadoc for 2.6.3 can be seen here. Notice, this class also uses Bean Validation constraints and will be honored by the JSF layer. Finally, create "MovieSessionBean" stateless EJB with all the business logic such that it looks like: package org.glassfish.samples;import com.mongodb.BasicDBObject;import com.mongodb.DB;import com.mongodb.DBCollection;import com.mongodb.DBCursor;import com.mongodb.DBObject;import com.mongodb.Mongo;import java.net.UnknownHostException;import java.util.ArrayList;import java.util.List;import javax.annotation.PostConstruct;import javax.ejb.Stateless;import javax.inject.Inject;import javax.inject.Named;/** * @author arun */@Stateless@Namedpublic class MovieSessionBean { @Inject Movie movie; DBCollection movieColl; @PostConstruct private void initDB() throws UnknownHostException { Mongo m = new Mongo(); DB db = m.getDB("movieDB"); movieColl = db.getCollection("movies"); if (movieColl == null) { movieColl = db.createCollection("movies", null); } } public void createMovie() { BasicDBObject doc = movie.toDBObject(); movieColl.insert(doc); } public List<Movie> getMovies() { List<Movie> movies = new ArrayList(); DBCursor cur = movieColl.find(); System.out.println("getMovies: Found " + cur.size() + " movie(s)"); for (DBObject dbo : cur.toArray()) { movies.add(Movie.fromDBObject(dbo)); } return movies; }} The database is initialized in @PostConstruct. Instead of a working with a database table, NoSQL databases work with a schema-less document. The "Movie" class is the document in our case and stored in the collection "movies". The collection allows us to perform query functions on all movies. The "getMovies" method invokes "find" method on the collection which is equivalent to the SQL query "select * from movies" and then returns a List<Movie>. Also notice that there is no "persistence.xml" in the project. Right-click and run the project to see the output as: Enter some values in the text box and click on enter to see the result as: If you reached here then you've successfully used MongoDB in your Java EE 6 application, congratulations! Some food for thought and further play ... SQL to MongoDB mapping shows mapping between traditional SQL -> Mongo query language. Tutorial shows fun things you can do with MongoDB. Try the interactive online shell  The cookbook provides common ways of using MongoDB In terms of this project, here are some tasks that can be tried: Encapsulate database management in a JPA persistence provider. Is it even worth it because the capabilities are going to be very different ? MongoDB uses "BSonObject" class for JSON representation, add @XmlRootElement on a POJO and how a compatible JSON representation can be generated. This will make the fromXXX and toXXX methods redundant.

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  • Cryptographic Validation Explained

    - by MarkPearl
    We have been using LogicNP’s CryptoLicensing for some of our software and I was battling to understand how exactly the whole process worked. I was sent the following document which really helped explain it – so if you ever use the same tool it is well worth a read. Licensing Basics LogicNP CryptoLicensing For .Net is the most advanced and state-of-the art licensing and copy protection system you can use for your software. LogicNP CryptoLicensing System uses the latest cryptographic technology to generate and validate licenses. The cryptographic algorithm used is the RSA algorithm which consists of a pair of keys called as the generation key and the validation key. Data encrypted using the generation key can only be decrypted using the corresponding validation key. How does cryptographic validation work? When a new license project is created, a unique validation-generation key pair is created for the project. When LogicNP CryptoLicensing For .Net generates licenses, it encrypts the license settings using the generation key. The validation key can be safely distributed with your software and is used during validation. During license validation, LogicNP CryptoLicensing For .Net attempts to decrypt the encrypted license code using the validation key. If the decryption is successful, this means that the data was encrypted using the generation key, since only the corresponding validation key can decrypt data encrypted with the generation key. This further means that not only is the license valid but that it was generated by you and only you since nobody else has access to the generation key. Generation Key This key is used by CryptoLicensing Generator to generate encrypted license codes. This key is stored in the license project file, so the license project file must be kept secure and confidential and must be accorded the same care as any other critical asset such as source code. Validation Key This key is used for validating generated license codes. It is the same key displayed in the 'Get Validation Key And Code' dialog (Ctrl+K) and is used by your software when validating license codes (using LogicNP.CryptoLicensing.dll). Unlike the generation key, it is not necessary to keep this key secure and confidential. Note that the generation key pair is stored in the project file created by LogicNP CryptoLicensing For .Net, so it is very important to backup this file and to keep it secure. Once the file is lost, it is not possible to retrieve the key pair. FAQ Do I use the same validation key to validate all license codes? Yes, the validation key (and generation key) for the project remains the same; you use the same key to validate all license codes generated using the project. You can retrieve the validation key using the "Project" menu --> "Get Validation Key & Code" menu item. Can license codes generated using generation key from one project be validated using validation key of another project? No! Q. Is every generated license code unique? A. Yes, every license code generated by CryptoLicensing is guaranteed to be unique, even if you generate thousands of codes at a time. Q. What makes CryptoLicensing so secure? A. CryptoLicensing uses the latest cryptographic technology to generate and validate licenses. The cryptographic algorithm used is the RSA asymmetric key algorithm which can use upto 3072-bit keys. Given current computing power, it takes years to break a 3072-bit key. Q. Is is possible for a hacker to develop a keygen for my software? A. Impossible. The cryptographic algorithm used by CryptoLicensing consists of a pair of keys called as the generation key and the validation key. Data encrypted with one key can only be decrypted by the other key and vice versa. Licenses are generated using the generation key and validated using the validation key. Without the generation key, it is impossible to generate valid licenses. Q. What is the difference between validation key and generation key? Generation Key This key is used by CryptoLicensing Generator to generate encrypted license codes. This key is stored in the license project file, so the license project file must be kept secure and confidential and must be accorded the same care as any other critical asset such as source code. Validation Key This key is used for validating generated license codes. It is the same key displayed in the 'Get Validation Key And Code' dialog (Ctrl+K) and is used by your software when validating license codes (using LogicNP.CryptoLicensing.dll). Unlike the generation key, it is not necessary to keep this key secure and confidential. Q. Do I have to include the license project file (.licproj) with my software? A. No!!! This goes against the very essence of the security of the asymmetric cryptographic scheme because the project file contains both the validation and generation key. With your software, you only need to include the validation key which will be used to validate licenses generated by CryptoLicensing using the generation key. The license project file should be treated as any other valuable and confidential asset such as your source code. Q. Does the license service need the license project file? A. Yes. The license project file is needed whenever new licenses are generated (via the UI, via the API or via the license service). As just one example, the license service generates new machine-locked licenses when activated licenses are presented to it for activation, therefore the license service needs the license project file. Q. Is it possible to embed my own data in the generated licenses? A. Yes. You can embed any amount of additional data in the licenses. This data will have the same amount of security as the license code itself and will be tamper-proof. The embedded user data can be retrieved from your software. Q. What additional steps can I take to ensure that my software does not get cracked? A. There are many methods and techniques which can make it extremely difficult for a hacker to crack your software. See Writing Effective License Checking Code And Designing Effective Licenses for more information. Q. Why is the license service not working? A. The most common cause is not setting the CryptoLicense.LicenseServiceURL property before trying to validate a license. Make sure that this property is set to the correct URL where your license service is hosted. The most common cause after this is that the license project file on the web server where your license service is hosted is not the latest. This happens if you make changes to the license project (for example, set the 'Enable With Serials' setting for a profile), but don't upload the updated project file to your web server. Q. Why are my serials not working? Serial codes require the user of a license service. See Using Serial Codes for more details. Also see the earlier question 'Why is the license service not working?' Q. Is the same validation key used to validate license codes generated from different profiles. A. Yes. Profiles are just pre specified license settings for quickly generating licenses having those settings. The actual license code is still generated using the license project's cryptographic generation key and thus, can be validated using the project's validation key. Q. Why are changes made to a profile not getting saved? A. Simply changing license settings via UI and saving the license project does not save those license settings to the active profile. You must first save the license settings to a profile using the Save/Save As command from the Profiles menu (see above). Q. Why is validation of activated licenses failing from CryptoLicensing Generator, but works from my software? A. Make sure that you have specified the URL of the license service using the Project Properties Dialog. Also see the earlier question 'Why is the license service not working?' Q. How can I extend the trial period of my customer? A. To extend the evaluation period of the customer, simply send him a new license code specifying the desired evaluation limits. Evaluation information such as the current used days, executions, etc are stored in garbled form in a registry location which is derived from the license code. Therefore, when a new license code is used, the old evaluation information will not be used and a new evaluation period will be started.

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