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  • Accessing py2exe program over network in Windows 98 throws ImportErrors

    - by darvids0n
    I'm running a py2exe-compiled python program from one server machine on a number of client machines (mapped to a network drive on every machine, say W:). For Windows XP and later machines, have so far had zero problems with Python picking up W:\python23.dll (yes, I'm using Python 2.3.5 for W98 compatibility and all that). It will then use W:\zlib.pyd to decompress W:\library.zip containing all the .pyc files like os and such, which are then imported and the program runs no problems. The issue I'm getting is on some Windows 98 SE machines (note: SOME Windows 98 SE machines, others seem to work with no apparent issues). What happens is, the program runs from W:, the W:\python23.dll is, I assume, found (since I'm getting Python ImportErrors, we'd need to be able to execute a Python import statement), but a couple of things don't work: 1) If W:\library.zip contains the only copy of the .pyc files, I get ZipImportError: can't decompress data; zlib not available (nonsense, considering W:\zlib.pyd IS available and works fine with the XP and higher machines on the same network). 2) If the .pyc files are actually bundled INSIDE the python exe by py2exe, OR put in the same directory as the .exe, OR put into a named subdirectory which is then set as part of the PYTHONPATH variable (e.g W:\pylib), I get ImportError: no module named os (os is the first module imported, before sys and anything else). Come to think of it, sys.path wouldn't be available to search if os was imported before it maybe? I'll try switching the order of those imports but my question still stands: Why is this a sporadic issue, working on some networks but not on others? And how would I force Python to find the files that are bundled inside the very executable I run? I have immediate access to the working Windows 98 SE machine, but I only get access to the non-working one (a customer of mine) every morning before their store opens. Thanks in advance! EDIT: Okay, big step forward. After debugging with PY2EXE_VERBOSE, the problem occurring on the specific W98SE machine is that it's not using the right path syntax when looking for imports. Firstly, it doesn't seem to read the PYTHONPATH environment variable (there may be a py2exe-specific one I'm not aware of, like PY2EXE_VERBOSE). Secondly, it only looks in one place before giving up (if the files are bundled inside the EXE, it looks there. If not, it looks in library.zip). EDIT 2: In fact, according to this, there is a difference between the sys.path in the Python interpreter and that of Py2exe executables. Specifically, sys.path contains only a single entry: the full pathname of the shared code archive. Blah. No fallbacks? Not even the current working directory? I'd try adding W:\ to PATH, but py2exe doesn't conform to any sort of standards for locating system libraries, so it won't work. Now for the interesting bit. The path it tries to load atexit, os, etc. from is: W:\\library.zip\<module>.<ext> Note the single slash after library.zip, but the double slash after the drive letter (someone correct me if this is intended and should work). It looks like if this is a string literal, then since the slash isn't doubled, it's read as an (invalid) escape sequence and the raw character is printed (giving W:\library.zipos.pyd, W:\library.zipos.dll, ... instead of with a slash); if it is NOT a string literal, the double slash might not be normpath'd automatically (as it should be) and so the double slash confuses the module loader. Like I said, I can't just set PYTHONPATH=W:\\library.zip\\ because it ignores that variable. It may be worth using sys.path.append at the start of my program but hard-coding module paths is an absolute LAST resort, especially since the problem occurs in ONE configuration of an outdated OS. Any ideas? I have one, which is to normpath the sys.path.. pity I need os for that. Another is to just append os.getenv('PATH') or os.getenv('PYTHONPATH') to sys.path... again, needing the os module. The site module also fails to initialise, so I can't use a .pth file. I also recently tried the following code at the start of the program: for pth in sys.path: fErr.write(pth) fErr.write(' to ') pth.replace('\\\\','\\') # Fix Windows 98 pathing issues fErr.write(pth) fErr.write('\n') But it can't load linecache.pyc, or anything else for that matter; it can't actually execute those commands from the looks of things. Is there any way to use built-in functionality which doesn't need linecache to modify the sys.path dynamically? Or am I reduced to hard-coding the correct sys.path?

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  • Mutating the expression tree of a predicate to target another type

    - by Jon
    Intro In the application I 'm currently working on, there are two kinds of each business object: the "ActiveRecord" type, and the "DataContract" type. So for example, we have: namespace ActiveRecord { class Widget { public int Id { get; set; } } } namespace DataContracts { class Widget { public int Id { get; set; } } } The database access layer takes care of "translating" between hierarchies: you can tell it to update a DataContracts.Widget, and it will magically create an ActiveRecord.Widget with the same property values and save that. The problem I have surfaced when attempting to refactor this database access layer. The Problem I want to add methods like the following to the database access layer: // Widget is DataContract.Widget interface DbAccessLayer { IEnumerable<Widget> GetMany(Expression<Func<Widget, bool>> predicate); } The above is a simple general-use "get" method with custom predicate. The only point of interest is that I 'm not passing in an anonymous function but rather an expression tree. This is done because inside DbAccessLayer we have to query ActiveRecord.Widget efficiently (LINQ to SQL) and not have the database return all ActiveRecord.Widget instances and then filter the enumerable collection. We need to pass in an expression tree, so we ask for one as the parameter for GetMany. The snag: the parameter we have needs to be magically transformed from an Expression<Func<DataContract.Widget, bool>> to an Expression<Func<ActiveRecord.Widget, bool>>. This is where I haven't managed to pull it off... Attempted Solution What we 'd like to do inside GetMany is: IEnumerable<DataContract.Widget> GetMany( Expression<Func<DataContract.Widget, bool>> predicate) { var lambda = Expression.Lambda<Func<ActiveRecord.Widget, bool>>( predicate.Body, predicate.Parameters); // use lambda to query ActiveRecord.Widget and return some value } This won't work because in a typical scenario, for example if: predicate == w => w.Id == 0; ...the expression tree contains a MemberAccessExpression instance which has a MemberInfo property (named Member) that point to members of DataContract.Widget. There are also ParameterExpression instances both in the expression tree and in its parameter expression collection (predicate.Parameters); After searching a bit, I found System.Linq.Expressions.ExpressionVisitor (its source can be found here in the context of a how-to, very helpful) which is a convenient way to modify an expression tree. Armed with this, I implemented a visitor. This simple visitor only takes care of changing the types in member access and parameter expressions. It may not be complete, but it's fine for the expression w => w.Id == 0. internal class Visitor : ExpressionVisitor { private readonly Func<Type, Type> dataContractToActiveRecordTypeConverter; public Visitor(Func<Type, Type> dataContractToActiveRecordTypeConverter) { this.dataContractToActiveRecordTypeConverter = dataContractToActiveRecordTypeConverter; } protected override Expression VisitMember(MemberExpression node) { var dataContractType = node.Member.ReflectedType; var activeRecordType = this.dataContractToActiveRecordTypeConverter(dataContractType); var converted = Expression.MakeMemberAccess( base.Visit(node.Expression), activeRecordType.GetProperty(node.Member.Name)); return converted; } protected override Expression VisitParameter(ParameterExpression node) { var dataContractType = node.Type; var activeRecordType = this.dataContractToActiveRecordTypeConverter(dataContractType); return Expression.Parameter(activeRecordType, node.Name); } } With this visitor, GetMany becomes: IEnumerable<DataContract.Widget> GetMany( Expression<Func<DataContract.Widget, bool>> predicate) { var visitor = new Visitor(...); var lambda = Expression.Lambda<Func<ActiveRecord.Widget, bool>>( visitor.Visit(predicate.Body), predicate.Parameters.Select(p => visitor.Visit(p)); var widgets = ActiveRecord.Widget.Repository().Where(lambda); // This is just for reference, see below Expression<Func<ActiveRecord.Widget, bool>> referenceLambda = w => w.Id == 0; // Here we 'd convert the widgets to instances of DataContract.Widget and // return them -- this has nothing to do with the question though. } Results The good news is that lambda is constructed just fine. The bad news is that it isn't working; it's blowing up on me when I try to use it (the exception messages are really not helpful at all). I have examined the lambda my code produces and a hardcoded lambda with the same expression; they look exactly the same. I spent hours in the debugger trying to find some difference, but I can't. When predicate is w => w.Id == 0, lambda looks exactly like referenceLambda. But the latter works with e.g. IQueryable<T>.Where, while the former does not (I have tried this in the immediate window of the debugger). I should also mention that when predicate is w => true, it all works just fine. Therefore I am assuming that I 'm not doing enough work in Visitor, but I can't find any more leads to follow on. Can someone point me in the right direction? Thanks in advance for your help!

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  • Neural Network Always Produces Same/Similar Outputs for Any Input

    - by l33tnerd
    I have a problem where I am trying to create a neural network for Tic-Tac-Toe. However, for some reason, training the neural network causes it to produce nearly the same output for any given input. I did take a look at Artificial neural networks benchmark, but my network implementation is built for neurons with the same activation function for each neuron, i.e. no constant neurons. To make sure the problem wasn't just due to my choice of training set (1218 board states and moves generated by a genetic algorithm), I tried to train the network to reproduce XOR. The logistic activation function was used. Instead of using the derivative, I multiplied the error by output*(1-output) as some sources suggested that this was equivalent to using the derivative. I can put the Haskell source on HPaste, but it's a little embarrassing to look at. The network has 3 layers: the first layer has 2 inputs and 4 outputs, the second has 4 inputs and 1 output, and the third has 1 output. Increasing to 4 neurons in the second layer didn't help, and neither did increasing to 8 outputs in the first layer. I then calculated errors, network output, bias updates, and the weight updates by hand based on http://hebb.mit.edu/courses/9.641/2002/lectures/lecture04.pdf to make sure there wasn't an error in those parts of the code (there wasn't, but I will probably do it again just to make sure). Because I am using batch training, I did not multiply by x in equation (4) there. I am adding the weight change, though http://www.faqs.org/faqs/ai-faq/neural-nets/part2/section-2.html suggests to subtract it instead. The problem persisted, even in this simplified network. For example, these are the results after 500 epochs of batch training and of incremental training. Input |Target|Output (Batch) |Output(Incremental) [1.0,1.0]|[0.0] |[0.5003781562785173]|[0.5009731800870864] [1.0,0.0]|[1.0] |[0.5003740346965251]|[0.5006347214672715] [0.0,1.0]|[1.0] |[0.5003734471544522]|[0.500589332376345] [0.0,0.0]|[0.0] |[0.5003674110937019]|[0.500095157458231] Subtracting instead of adding produces the same problem, except everything is 0.99 something instead of 0.50 something. 5000 epochs produces the same result, except the batch-trained network returns exactly 0.5 for each case. (Heck, even 10,000 epochs didn't work for batch training.) Is there anything in general that could produce this behavior? Also, I looked at the intermediate errors for incremental training, and the although the inputs of the hidden/input layers varied, the error for the output neuron was always +/-0.12. For batch training, the errors were increasing, but extremely slowly and the errors were all extremely small (x10^-7). Different initial random weights and biases made no difference, either. Note that this is a school project, so hints/guides would be more helpful. Although reinventing the wheel and making my own network (in a language I don't know well!) was a horrible idea, I felt it would be more appropriate for a school project (so I know what's going on...in theory, at least. There doesn't seem to be a computer science teacher at my school). EDIT: Two layers, an input layer of 2 inputs to 8 outputs, and an output layer of 8 inputs to 1 output, produces much the same results: 0.5+/-0.2 (or so) for each training case. I'm also playing around with pyBrain, seeing if any network structure there will work. Edit 2: I am using a learning rate of 0.1. Sorry for forgetting about that. Edit 3: Pybrain's "trainUntilConvergence" doesn't get me a fully trained network, either, but 20000 epochs does, with 16 neurons in the hidden layer. 10000 epochs and 4 neurons, not so much, but close. So, in Haskell, with the input layer having 2 inputs & 2 outputs, hidden layer with 2 inputs and 8 outputs, and output layer with 8 inputs and 1 output...I get the same problem with 10000 epochs. And with 20000 epochs. Edit 4: I ran the network by hand again based on the MIT PDF above, and the values match, so the code should be correct unless I am misunderstanding those equations. Some of my source code is at http://hpaste.org/42453/neural_network__not_working; I'm working on cleaning my code somewhat and putting it in a Github (rather than a private Bitbucket) repository. All of the relevant source code is now at https://github.com/l33tnerd/hsann.

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  • cant populate cells with an array when i have loaded a second UITableViewController

    - by richard Stephenson
    hi there, im very new to iphone programming, im creating my first app, (a world cup one) the first view is a table view. the cell text label is filled with an array, so it shows all the groups (group a, B, c,ect) then when you select a group, it pulls on another UITableViewcontroller, but whatever i do i cant set the text label of the cells (e.g france,mexico,south africa, etc. infact nothin i do to the cellForRowAtIndexPath makes a difference , could someone tell me what im doing wrong please Thanks `here is my code for the view controller #import "GroupADetailViewController.h" @implementation GroupADetailViewController @synthesize groupLabel = _groupLabel; @synthesize groupADetail = _groupADetail; @synthesize teamsInGroupA; #pragma mark Memory management - (void)dealloc { [_groupADetail release]; [_groupLabel release]; [super dealloc]; } #pragma mark View lifecycle - (void)viewDidLoad { [super viewDidLoad]; // Set the number label to show the number data teamsInGroupA = [[NSArray alloc]initWithObjects:@"France",@"Mexico",@"Uruguay",@"South Africa",nil]; NSLog(@"loaded"); // Set the title to also show the number data [[self navigationItem]setTitle:@"Group A"]; //[[self navigationItem]cell.textLabel.text:@"test"]; //[[self navigationItem] setTitle[NSString String } - (void)viewDidUnload { [self setgroupLabel:nil]; } #pragma mark Table view methods - (NSInteger)numberOfSectionsInTableView:(UITableView*)tableView { // Return the number of sections in the table view return 1; } - (NSInteger)tableView:(UITableView*)tableView numberOfRowsInSection:(NSInteger)section { // Return the number of rows in a specific section // Since we only have one section, just return the number of rows in the table return 4; NSLog:("count is %d",[teamsInGroupA count]); } - (UITableViewCell*)tableView:(UITableView*)tableView cellForRowAtIndexPath:(NSIndexPath*)indexPath { static NSString *cellIdentifier2 = @"Cell2"; // Reuse an existing cell if one is available for reuse UITableViewCell *cell = [tableView dequeueReusableCellWithIdentifier:cellIdentifier2]; // If no cell was available, create a new one if (cell == nil) { NSLog(@"no cell, creating"); cell = [[[UITableViewCell alloc] initWithStyle:UITableViewCellStyleDefault reuseIdentifier:cellIdentifier2] autorelease]; [cell setAccessoryType:UITableViewCellAccessoryDisclosureIndicator]; } NSLog(@"cell already there"); // Configure the cell to show the data for this row //[[cell textLabel]setText:[NSString string //[[cell textLabel]setText:[teamsInGroupA objectAtIndex:indexPath.row]]; //NSUInteger row = [indexPath row]; //[cell setText:[[teamsInGroupA objectAtIndex:indexPath:row]retain]]; //cell.textLabel.text:@"Test" [[cell textLabel]setText:[teamsInGroupA objectAtIndex:indexPath.row]]; return cell; } @end #import "GroupADetailViewController.h" @implementation GroupADetailViewController @synthesize groupLabel = _groupLabel; @synthesize groupADetail = _groupADetail; @synthesize teamsInGroupA; #pragma mark Memory management - (void)dealloc { [_groupADetail release]; [_groupLabel release]; [super dealloc]; } #pragma mark View lifecycle - (void)viewDidLoad { [super viewDidLoad]; // Set the number label to show the number data teamsInGroupA = [[NSArray alloc]initWithObjects:@"France",@"Mexico",@"Uruguay",@"South Africa",nil]; NSLog(@"loaded"); // Set the title to also show the number data [[self navigationItem]setTitle:@"Group A"]; //[[self navigationItem]cell.textLabel.text:@"test"]; //[[self navigationItem] setTitle[NSString String } - (void)viewDidUnload { [self setgroupLabel:nil]; } #pragma mark Table view methods - (NSInteger)numberOfSectionsInTableView:(UITableView*)tableView { // Return the number of sections in the table view return 1; } - (NSInteger)tableView:(UITableView*)tableView numberOfRowsInSection:(NSInteger)section { // Return the number of rows in a specific section // Since we only have one section, just return the number of rows in the table return 4; NSLog:("count is %d",[teamsInGroupA count]); } - (UITableViewCell*)tableView:(UITableView*)tableView cellForRowAtIndexPath:(NSIndexPath*)indexPath { static NSString *cellIdentifier2 = @"Cell2"; // Reuse an existing cell if one is available for reuse UITableViewCell *cell = [tableView dequeueReusableCellWithIdentifier:cellIdentifier2]; // If no cell was available, create a new one if (cell == nil) { NSLog(@"no cell, creating"); cell = [[[UITableViewCell alloc] initWithStyle:UITableViewCellStyleDefault reuseIdentifier:cellIdentifier2] autorelease]; [cell setAccessoryType:UITableViewCellAccessoryDisclosureIndicator]; } NSLog(@"cell already there"); // Configure the cell to show the data for this row //[[cell textLabel]setText:[NSString string //[[cell textLabel]setText:[teamsInGroupA objectAtIndex:indexPath.row]]; //NSUInteger row = [indexPath row]; //[cell setText:[[teamsInGroupA objectAtIndex:indexPath:row]retain]]; //cell.textLabel.text:@"Test" [[cell textLabel]setText:[teamsInGroupA objectAtIndex:indexPath.row]]; return cell; } @end

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  • Performance of delegate and method group

    - by BlueFox
    Hi I was investigating the performance hit of creating Cachedependency objects, so I wrote a very simple test program as follows: using System; using System.Collections.Generic; using System.Diagnostics; using System.Web.Caching; namespace Test { internal class Program { private static readonly string[] keys = new[] {"Abc"}; private static readonly int MaxIteration = 10000000; private static void Main(string[] args) { Debug.Print("first set"); test7(); test6(); test5(); test4(); test3(); test2(); Debug.Print("second set"); test2(); test3(); test4(); test5(); test6(); test7(); } private static void test2() { DateTime start = DateTime.Now; var list = new List<CacheDependency>(); for (int i = 0; i < MaxIteration; i++) { list.Add(new CacheDependency(null, keys)); } Debug.Print("test2 Time: " + (DateTime.Now - start)); } private static void test3() { DateTime start = DateTime.Now; var list = new List<Func<CacheDependency>>(); for (int i = 0; i < MaxIteration; i++) { list.Add(() => new CacheDependency(null, keys)); } Debug.Print("test3 Time: " + (DateTime.Now - start)); } private static void test4() { var p = new Program(); DateTime start = DateTime.Now; var list = new List<Func<CacheDependency>>(); for (int i = 0; i < MaxIteration; i++) { list.Add(p.GetDep); } Debug.Print("test4 Time: " + (DateTime.Now - start)); } private static void test5() { var p = new Program(); DateTime start = DateTime.Now; var list = new List<Func<CacheDependency>>(); for (int i = 0; i < MaxIteration; i++) { list.Add(() => { return p.GetDep(); }); } Debug.Print("test5 Time: " + (DateTime.Now - start)); } private static void test6() { DateTime start = DateTime.Now; var list = new List<Func<CacheDependency>>(); for (int i = 0; i < MaxIteration; i++) { list.Add(GetDepSatic); } Debug.Print("test6 Time: " + (DateTime.Now - start)); } private static void test7() { DateTime start = DateTime.Now; var list = new List<Func<CacheDependency>>(); for (int i = 0; i < MaxIteration; i++) { list.Add(() => { return GetDepSatic(); }); } Debug.Print("test7 Time: " + (DateTime.Now - start)); } private CacheDependency GetDep() { return new CacheDependency(null, keys); } private static CacheDependency GetDepSatic() { return new CacheDependency(null, keys); } } } But I can't understand why these result looks like this: first set test7 Time: 00:00:00.4840277 test6 Time: 00:00:02.2041261 test5 Time: 00:00:00.1910109 test4 Time: 00:00:03.1401796 test3 Time: 00:00:00.1820105 test2 Time: 00:00:08.5394884 second set test2 Time: 00:00:07.7324423 test3 Time: 00:00:00.1830105 test4 Time: 00:00:02.3561347 test5 Time: 00:00:00.1750100 test6 Time: 00:00:03.2941884 test7 Time: 00:00:00.1850106 In particular: 1. Why is test4 and test6 much slower than their delegate version? I also noticed that Resharper specifically has a comment on the delegate version suggesting change test5 and test7 to "Covert to method group". Which is the same as test4 and test6 but they're actually slower? 2. I don't seem a consistent performance difference when calling test4 and test6, shouldn't static calls to be always faster?

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  • Why is insertion into my tree faster on sorted input than random input?

    - by Juliet
    Now I've always heard binary search trees are faster to build from randomly selected data than ordered data, simply because ordered data requires explicit rebalancing to keep the tree height at a minimum. Recently I implemented an immutable treap, a special kind of binary search tree which uses randomization to keep itself relatively balanced. In contrast to what I expected, I found I can consistently build a treap about 2x faster and generally better balanced from ordered data than unordered data -- and I have no idea why. Here's my treap implementation: http://pastebin.com/VAfSJRwZ And here's a test program: using System; using System.Collections.Generic; using System.Linq; using System.Diagnostics; namespace ConsoleApplication1 { class Program { static Random rnd = new Random(); const int ITERATION_COUNT = 20; static void Main(string[] args) { List<double> rndTimes = new List<double>(); List<double> orderedTimes = new List<double>(); rndTimes.Add(TimeIt(50, RandomInsert)); rndTimes.Add(TimeIt(100, RandomInsert)); rndTimes.Add(TimeIt(200, RandomInsert)); rndTimes.Add(TimeIt(400, RandomInsert)); rndTimes.Add(TimeIt(800, RandomInsert)); rndTimes.Add(TimeIt(1000, RandomInsert)); rndTimes.Add(TimeIt(2000, RandomInsert)); rndTimes.Add(TimeIt(4000, RandomInsert)); rndTimes.Add(TimeIt(8000, RandomInsert)); rndTimes.Add(TimeIt(16000, RandomInsert)); rndTimes.Add(TimeIt(32000, RandomInsert)); rndTimes.Add(TimeIt(64000, RandomInsert)); rndTimes.Add(TimeIt(128000, RandomInsert)); string rndTimesAsString = string.Join("\n", rndTimes.Select(x => x.ToString()).ToArray()); orderedTimes.Add(TimeIt(50, OrderedInsert)); orderedTimes.Add(TimeIt(100, OrderedInsert)); orderedTimes.Add(TimeIt(200, OrderedInsert)); orderedTimes.Add(TimeIt(400, OrderedInsert)); orderedTimes.Add(TimeIt(800, OrderedInsert)); orderedTimes.Add(TimeIt(1000, OrderedInsert)); orderedTimes.Add(TimeIt(2000, OrderedInsert)); orderedTimes.Add(TimeIt(4000, OrderedInsert)); orderedTimes.Add(TimeIt(8000, OrderedInsert)); orderedTimes.Add(TimeIt(16000, OrderedInsert)); orderedTimes.Add(TimeIt(32000, OrderedInsert)); orderedTimes.Add(TimeIt(64000, OrderedInsert)); orderedTimes.Add(TimeIt(128000, OrderedInsert)); string orderedTimesAsString = string.Join("\n", orderedTimes.Select(x => x.ToString()).ToArray()); Console.WriteLine("Done"); } static double TimeIt(int insertCount, Action<int> f) { Console.WriteLine("TimeIt({0}, {1})", insertCount, f.Method.Name); List<double> times = new List<double>(); for (int i = 0; i < ITERATION_COUNT; i++) { Stopwatch sw = Stopwatch.StartNew(); f(insertCount); sw.Stop(); times.Add(sw.Elapsed.TotalMilliseconds); } return times.Average(); } static void RandomInsert(int insertCount) { Treap<double> tree = new Treap<double>((x, y) => x.CompareTo(y)); for (int i = 0; i < insertCount; i++) { tree = tree.Insert(rnd.NextDouble()); } } static void OrderedInsert(int insertCount) { Treap<double> tree = new Treap<double>((x, y) => x.CompareTo(y)); for(int i = 0; i < insertCount; i++) { tree = tree.Insert(i + rnd.NextDouble()); } } } } And here's a chart comparing random and ordered insertion times in milliseconds: Insertions Random Ordered RandomTime / OrderedTime 50 1.031665 0.261585 3.94 100 0.544345 1.377155 0.4 200 1.268320 0.734570 1.73 400 2.765555 1.639150 1.69 800 6.089700 3.558350 1.71 1000 7.855150 4.704190 1.67 2000 17.852000 12.554065 1.42 4000 40.157340 22.474445 1.79 8000 88.375430 48.364265 1.83 16000 197.524000 109.082200 1.81 32000 459.277050 238.154405 1.93 64000 1055.508875 512.020310 2.06 128000 2481.694230 1107.980425 2.24 I don't see anything in the code which makes ordered input asymptotically faster than unordered input, so I'm at a loss to explain the difference. Why is it so much faster to build a treap from ordered input than random input?

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  • Embedding mercurial revision information in Visual Studio c# projects automatically

    - by Mark Booth
    Original Problem In building our projects, I want the mercurial id of each repository to be embedded within the product(s) of that repository (the library, application or test application). I find it makes it so much easier to debug an application ebing run by custiomers 8 timezones away if you know precisely what went into building the particular version of the application they are using. As such, every project (application or library) in our systems implement a way of getting at the associated revision information. I also find it very useful to be able to see if an application has been compiled with clean (un-modified) changesets from the repository. 'Hg id' usefully appends a + to the changeset id when there are uncommitted changes in a repository, so this allows is to easily see if people are running a clean or a modified version of the code. My current solution is detailed below, and fulfills the basic requirements, but there are a number of problems with it. Current Solution At the moment, to each and every Visual Studio solution, I add the following "Pre-build event command line" commands: cd $(ProjectDir) HgID I also add an HgID.bat file to the Project directory: @echo off type HgId.pre > HgId.cs For /F "delims=" %%a in ('hg id') Do <nul >>HgID.cs set /p = @"%%a" echo ; >> HgId.cs echo } >> HgId.cs echo } >> HgId.cs along with an HgId.pre file, which is defined as: namespace My.Namespace { /// <summary> Auto generated Mercurial ID class. </summary> internal class HgID { /// <summary> Mercurial version ID [+ is modified] [Named branch]</summary> public const string Version = When I build my application, the pre-build event is triggered on all libraries, creating a new HgId.cs file (which is not kept under revision control) and causing the library to be re-compiled with with the new 'hg id' string in 'Version'. Problems with the current solution The main problem is that since the HgId.cs is re-created at each pre-build, every time we need to compile anything, all projects in the current solution are re-compiled. Since we want to be able to easily debug into our libraries, we usually keep many libraries referenced in our main application solution. This can result in build times which are significantly longer than I would like. Ideally I would like the libraries to compile only if the contents of the HgId.cs file has actually changed, as opposed to having been re-created with exactly the same contents. The second problem with this method is it's dependence on specific behaviour of the windows shell. I've already had to modify the batch file several times, since the original worked under XP but not Vista, the next version worked under Vista but not XP and finally I managed to make it work with both. Whether it will work with Windows 7 however is anyones guess and as time goes on, I see it more likely that contractors will expect to be able to build our apps on their Windows 7 boxen. Finally, I have an aesthetic problem with this solution, batch files and bodged together template files feel like the wrong way to do this. My actual questions How would you solve/how are you solving the problem I'm trying to solve? What better options are out there than what I'm currently doing? Rejected Solutions to these problems Before I implemented the current solution, I looked at Mercurials Keyword extension, since it seemed like the obvious solution. However the more I looked at it and read peoples opinions, the more that I came to the conclusion that it wasn't the right thing to do. I also remember the problems that keyword substitution has caused me in projects at previous companies (just the thought of ever having to use Source Safe again fills me with a feeling of dread *8'). Also, I don't particularly want to have to enable Mercurial extensions to get the build to complete. I want the solution to be self contained, so that it isn't easy for the application to be accidentally compiled without the embedded version information just because an extension isn't enabled or the right helper software hasn't been installed. I also thought of writing this in a better scripting language, one where I would only write HgId.cs file if the content had actually changed, but all of the options I could think of would require my co-workers, contractors and possibly customers to have to install software they might not otherwise want (for example cygwin). Any other options people can think of would be appreciated. Update Partial solution Having played around with it for a while, I've managed to get the HgId.bat file to only overwrite the HgId.cs file if it changes: @echo off type HgId.pre > HgId.cst For /F "delims=" %%a in ('hg id') Do <nul >>HgId.cst set /p = @"%%a" echo ; >> HgId.cst echo } >> HgId.cst echo } >> HgId.cst fc HgId.cs HgId.cst >NUL if %errorlevel%==0 goto :ok copy HgId.cst HgId.cs :ok del HgId.cst Problems with this solution Even though HgId.cs is no longer being re-created every time, Visual Studio still insists on compiling everything every time. I've tried looking for solutions and tried checking "Only build startup projects and dependencies on Run" in Tools|Options|Projects and Solutions|Build and Run but it makes no difference. The second problem also remains, and now I have no way to test if it will work with Vista, since that contractor is no longer with us. If anyone can test this batch file on a Windows 7 and/or Vista box, I would appreciate hearing how it went. Finally, my aesthetic problem with this solution, is even strnger than it was before, since the batch file is more complex and this there is now more to go wrong. If you can think of any better solution, I would love to hear about them.

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  • jQuery Cycle Plugin - Content not cycling

    - by fmz
    I am setting up a page with jQuery's Cycle plugin and have four divs set to fade. I have the code in place, the images set, but it doesn't cycle properly. Firefox says there is a problem with the following code: <script type="text/javascript"> $(document).ready(function() { $('.slideshow').cycle({ fx: 'fade' }); }); </script> Here is the html: <div class="slideshow"> <div id="mainImg-1" class="slide"> <div class="quote"> <h2>Building Big Relationships with Small Business.</h2> <p>&ldquo;This is quote Number One.<br /> They are there when I need them the most.&rdquo;</p> <p><span class="author">Jane Doe &ndash; Charlotte Flower Shop</span></p> <div class="help"><a href="cb_services.html">Let Us Help You</a></div> </div> </div> <div id="mainImg-2" class="slide"> <div class="quote"> <h2>Building Big Relationships with Small Business.</h2> <p>&ldquo;This is quote Number Two.<br /> They are there when I need them the most.&rdquo;</p> <p><span class="author">Jane Doe &ndash; Charlotte Flower Shop</span></p> <div class="help"><a href="cb_services.html">Let Us Help You</a></div> </div> </div> <div id="mainImg-3" class="slide"> <div class="quote"> <h2>Building Big Relationships with Small Business.</h2> <p>&ldquo;This is quote Number three.<br /> They are there when I need them the most.&rdquo;</p> <p><span class="author">Jane Doe &ndash; Charlotte Flower Shop</span></p> <div class="help"><a href="cb_services.html">Let Us Help You</a></div> </div> </div> <div id="mainImg-4" class="slide"> <div class="quote"> <h2>Building Big Relationships with Small Business.</h2> <p>&ldquo;This is quote Number Fout.<br /> They are there when I need them the most.&rdquo;</p> <p><span class="author">Jane Doe &ndash; Charlotte Flower Shop</span></p> <div class="help"><a href="cb_services.html">Let Us Help You</a></div> </div> </div> Here is the CSS: .slideshow { width: 946px; height: 283px; border: 1px solid #c29c5d; margin: 8px; overflow: hidden; z-index: 1; } #mainImg-1 { width: 946px; height: 283px; background: url(../_images/main.jpg) no-repeat 9px 9px; } #mainImg-2 { width: 946px; height: 283px; background: url(../_images/main.jpg) no-repeat 9px 9px; } #mainImg-3 { width: 946px; height: 283px; background: url(../_images/main.jpg) no-repeat 9px 9px; } #mainImg-4 { width: 946px; height: 283px; background: url(../_images/main.jpg) no-repeat 9px 9px; } #mainImg-1 .quote, #mainImg-2 .quote, #mainImg-3 .quote, #mainImg-4 .quote { width: 608px; height: 168px; float: right; margin: 80px 11px 0 0; background: url(../_images/bg_quoteBox.png) repeat-x; } Before you go off and say, "hey, those images are all the same". You are right, the images are all the same right now, but the text should be rotating as well and there is a slight difference there. In addition, the fade should still show up. Anyway, you can see the dev page here: http://173.201.163.213/projectpath/first_trust/index.html I would appreciate some help to get this cycling through as it should. Thanks!

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  • How should I delete a child object from within a parent's slot? Possibly boost::asio specific.

    - by kaliatech
    I have written a network server class that maintains a std::set of network clients. The network clients emit a signal to the network server on disconnect (via boost::bind). When a network client disconnects, the client instance needs to be removed from the Set and eventually deleted. I would think this is a common pattern, but I am having problems that might, or might not, be specific to ASIO. I've tried to trim down to just the relevant code: /** NetworkServer.hpp **/ class NetworkServices : private boost::noncopyable { public: NetworkServices(void); ~NetworkServices(void); private: void run(); void onNetworkClientEvent(NetworkClientEvent&); private: std::set<boost::shared_ptr<const NetworkClient>> clients; }; /** NetworkClient.cpp **/ void NetworkServices::run() { running = true; boost::asio::io_service::work work(io_service); //keeps service running even if no operations // This creates just one thread for the boost::asio async network services boost::thread iot(boost::bind(&NetworkServices::run_io_service, this)); while (running) { boost::system::error_code err; try { tcp::socket* socket = new tcp::socket(io_service); acceptor->accept(*socket, err); if (!err) { NetworkClient* networkClient = new NetworkClient(io_service, boost::shared_ptr<tcp::socket>(socket)); networkClient->networkClientEventSignal.connect(boost::bind(&NetworkServices::onNetworkClientEvent, this, _1)); clients.insert(boost::shared_ptr<NetworkClient>(networkClient)); networkClient->init(); //kicks off 1st asynch_read call } } // etc... } } void NetworkServices::onNetworkClientEvent(NetworkClientEvent& evt) { switch(evt.getType()) { case NetworkClientEvent::CLIENT_ERROR : { boost::shared_ptr<const NetworkClient> clientPtr = evt.getClient().getSharedPtr(); // ------ THIS IS THE MAGIC LINE ----- // If I keep this, the io_service hangs. If I comment it out, // everything works fine (but I never delete the disconnected NetworkClient). // If actually deleted the client here I might expect problems because it is the caller // of this method via boost::signal and bind. However, The clientPtr is a shared ptr, and a // reference is being kept in the client itself while signaling, so // I would the object is not going to be deleted from the heap here. That seems to be the case. // Never-the-less, this line makes all the difference, most likely because it controls whether or not the NetworkClient ever gets deleted. clients.erase(clientPtr); //I should probably put this socket clean-up in NetworkClient destructor. Regardless by doing this, // I would expect the ASIO socket stuff to be adequately cleaned-up after this. tcp::socket& socket = clientPtr->getSocket(); try { socket.shutdown(boost::asio::socket_base::shutdown_both); socket.close(); } catch(...) { CommServerContext::error("Error while shutting down and closing socket."); } break; } default : { break; } } } /** NetworkClient.hpp **/ class NetworkClient : public boost::enable_shared_from_this<NetworkClient>, Client { NetworkClient(boost::asio::io_service& io_service, boost::shared_ptr<tcp::socket> socket); virtual ~NetworkClient(void); inline boost::shared_ptr<const NetworkClient> getSharedPtr() const { return shared_from_this(); }; boost::signal <void (NetworkClientEvent&)> networkClientEventSignal; void onAsyncReadHeader(const boost::system::error_code& error, size_t bytes_transferred); }; /** NetworkClient.cpp - onAsyncReadHeader method called from io_service.run() thread as result of an async_read operation. Error condition usually result of an unexpected client disconnect.**/ void NetworkClient::onAsyncReadHeader( const boost::system::error_code& error, size_t bytes_transferred) { if (error) { //Make sure this instance doesn't get deleted from parent/slot deferencing //Alternatively, somehow schedule for future delete? boost::shared_ptr<const NetworkClient> clientPtr = getSharedPtr(); //Signal to service that this client is disconnecting NetworkClientEvent evt(*this, NetworkClientEvent::CLIENT_ERROR); networkClientEventSignal(evt); networkClientEventSignal.disconnect_all_slots(); return; } I believe it's not safe to delete the client from within the slot handler because the function return would be ... undefined? (Interestingly, it doesn't seem to blow up on me though.) So I've used boost:shared_ptr along with shared_from_this to make sure the client doesn't get deleted until all slots have been signaled. It doesn't seem to really matter though. I believe this question is not specific to ASIO, but the problem manifests in a peculiar way when using ASIO. I have one thread executing io_service.run(). All ASIO read/write operations are performed asynchronously. Everything works fine with multiple clients connecting/disconnecting UNLESS I delete my client object from the Set per the code above. If I delete my client object, the io_service seemingly deadlocks internally and no further asynchronous operations are performed unless I start another thread. I have try/catches around the io_service.run() call and have not been able to detect any errors. Questions: Are there best practices for deleting child objects, that are also signal emitters, from within parent slots? Any ideas as to why the io_service is hanging when I delete my network client object?

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  • App crashes when adding array data to table cells

    - by bassmandan
    I am trying to create a table view that loads a number of tweets into the table (one per cell etc). I am using NSXMLParser to get the information and have got as far as creating an array with the selection of tweets that I want. However, when I try to add them to the table cells, the app crashes on the line: cell.textLabel.text = cellValue; An NSLog before this shows in the console that the app is getting the correct data, so I am a bit stumped as to why this isn't working. This is the block of code that appears to be having the problem: - (UITableViewCell *)tableView:(UITableView *)tableView cellForRowAtIndexPath:(NSIndexPath *)indexPath { static NSString *CellIdentifier = @"Cell"; UITableViewCell *cell = [tableView dequeueReusableCellWithIdentifier:CellIdentifier]; if (cell == nil) { cell = [[UITableViewCell alloc] initWithStyle:UITableViewCellStyleDefault reuseIdentifier:CellIdentifier]; } // Set up the cell... NSString *cellValue = [statuses objectAtIndex:indexPath.row]; NSLog(@"%@", cellValue); cell.textLabel.text = cellValue; return cell;} If it makes a difference, I am using ARC and the latest version of XCode. I'm still quite new to all this, so if I need to give some extra information, let me know. Thanks in advance. Edit: Backtrace gives the following: * thread #1: tid = 0x2003, 0x918a19c6 libsystem_kernel.dylib`__pthread_kill + 10, stop reason = signal SIGABRT frame #0: 0x918a19c6 libsystem_kernel.dylib`__pthread_kill + 10 frame #1: 0x9968ff78 libsystem_c.dylib`pthread_kill + 106 frame #2: 0x99680bdd libsystem_c.dylib`abort + 167 frame #3: 0x03c93e78 libc++abi.dylib`_Unwind_DeleteException frame #4: 0x03c9189e libc++abi.dylib`_ZL17default_terminatev + 34 frame #5: 0x0154df4b libobjc.A.dylib`_objc_terminate + 94 frame #6: 0x03c918de libc++abi.dylib`_ZL19safe_handler_callerPFvvE + 13 frame #7: 0x03c91946 libc++abi.dylib`std::terminate() + 23 frame #8: 0x03c92ab2 libc++abi.dylib`__cxa_throw + 110 frame #9: 0x0154de15 libobjc.A.dylib`objc_exception_throw + 311 frame #10: 0x013bdced CoreFoundation`-[NSObject doesNotRecognizeSelector:] + 253 frame #11: 0x01322f00 CoreFoundation`___forwarding___ + 432 frame #12: 0x01322ce2 CoreFoundation`_CF_forwarding_prep_0 + 50 frame #13: 0x0015168f UIKit`-[UILabel setText:] + 56 frame #14: 0x00003088 Twitter`-[TwitterViewController tableView:cellForRowAtIndexPath:] + 376 at TwitterViewController.m:131 frame #15: 0x000ace0f UIKit`-[UITableView(UITableViewInternal) _createPreparedCellForGlobalRow:withIndexPath:] + 494 frame #16: 0x000ad589 UIKit`-[UITableView(UITableViewInternal) _createPreparedCellForGlobalRow:] + 69 frame #17: 0x00098dfd UIKit`-[UITableView(_UITableViewPrivate) _updateVisibleCellsNow:] + 1350 frame #18: 0x000a7851 UIKit`-[UITableView layoutSubviews] + 242 frame #19: 0x00052301 UIKit`-[UIView(CALayerDelegate) layoutSublayersOfLayer:] + 145 frame #20: 0x013bde72 CoreFoundation`-[NSObject performSelector:withObject:] + 66 frame #21: 0x01d6692d QuartzCore`-[CALayer layoutSublayers] + 266 frame #22: 0x01d70827 QuartzCore`CA::Layer::layout_if_needed(CA::Transaction*) + 231 frame #23: 0x01cf6fa7 QuartzCore`CA::Context::commit_transaction(CA::Transaction*) + 377 frame #24: 0x01cf8ea6 QuartzCore`CA::Transaction::commit() + 374 frame #25: 0x01d8430c QuartzCore`+[CATransaction flush] + 52 frame #26: 0x000124c6 UIKit`-[UIApplication _reportAppLaunchFinished] + 39 frame #27: 0x00012bd6 UIKit`-[UIApplication _runWithURL:payload:launchOrientation:statusBarStyle:statusBarHidden:] + 1324 frame #28: 0x00021743 UIKit`-[UIApplication handleEvent:withNewEvent:] + 1027 frame #29: 0x000221f8 UIKit`-[UIApplication sendEvent:] + 68 frame #30: 0x00015aa9 UIKit`_UIApplicationHandleEvent + 8196 frame #31: 0x012a6fa9 GraphicsServices`PurpleEventCallback + 1274 frame #32: 0x013901c5 CoreFoundation`__CFRUNLOOP_IS_CALLING_OUT_TO_A_SOURCE1_PERFORM_FUNCTION__ + 53 frame #33: 0x012f5022 CoreFoundation`__CFRunLoopDoSource1 + 146 frame #34: 0x012f390a CoreFoundation`__CFRunLoopRun + 2218 frame #35: 0x012f2db4 CoreFoundation`CFRunLoopRunSpecific + 212 frame #36: 0x012f2ccb CoreFoundation`CFRunLoopRunInMode + 123 frame #37: 0x000122a7 UIKit`-[UIApplication _run] + 576 frame #38: 0x00013a9b UIKit`UIApplicationMain + 1175 frame #39: 0x0000239d Twitter`main + 141 at main.m:16 frame #40: 0x00002305 Twitter`start + 53 Debugging console shows this: 2012-04-08 10:10:05.084 Twitter[25309:f803] ( { text = "Have you shared the Shakedown yet? http://t.co/WHrIC9w7"; }, { text = "For all you closet rocknrollas pencil in Sat 12th May The Rebirth of Rock n Roll Party. Haywire Saint @ The Good... http://t.co/OXHKlLIV"; }, { text = "4 weeks today: Vocal tracks will be getting recorded at The Premises Studios"; }, { text = "Rehearsal tonight in preparation to some big recording next month!"; }, { text = "haywire saint 'great taste.' Tune. \n\nhttp://t.co/GKmu5Lna http://t.co/0fii55Hw"; }, { text = "Meeting up with an old roadie for The Cure today. oh the stories...... http://t.co/UeUYccme"; }, { text = "Satisfying day of programming today.. Haywire Saint app coming along nicely with the custom music player ready to rock 'n' roll!"; }, { text = "Happy Friday Everyone!"; }, { text = "We had a great time at The Premises Studios yesterday. We'll be back there before long :D x"; }, { text = "I posted a new photo to Facebook http://t.co/73qAnCvk"; } ) 2012-04-08 10:10:05.093 Twitter[25309:f803] { text = "Have you shared the Shakedown yet? http://t.co/WHrIC9w7"; } 2012-04-08 10:10:05.094 Twitter[25309:f803] -[__NSCFDictionary isEqualToString:]: unrecognized selector sent to instance 0x6877a50 2012-04-08 10:10:05.096 Twitter[25309:f803] *** Terminating app due to uncaught exception 'NSInvalidArgumentException', reason: '-[__NSCFDictionary isEqualToString:]: unrecognized selector sent to instance 0x6877a50' *** First throw call stack: (0x13bc052 0x154dd0a 0x13bdced 0x1322f00 0x1322ce2 0x15168f 0x3088 0xace0f 0xad589 0x98dfd 0xa7851 0x52301 0x13bde72 0x1d6692d 0x1d70827 0x1cf6fa7 0x1cf8ea6 0x1d8430c 0x124c6 0x12bd6 0x21743 0x221f8 0x15aa9 0x12a6fa9 0x13901c5 0x12f5022 0x12f390a 0x12f2db4 0x12f2ccb 0x122a7 0x13a9b 0x239d 0x2305) terminate called throwing an exception2012-04-08 10:10:05.924 Twitter[25309:f803] -[__NSCFConstantString count]: unrecognized selector sent to instance 0x5b30

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  • c++ and c# speed compared

    - by Mack
    I was worried about C#'s speed when it deals with heavy calculations, when you need to use raw CPU power. I always thought that C++ is much faster than C# when it comes to calculations. So I did some quick tests. The first test computes prime numbers < an integer n, the second test computes some pandigital numbers. The idea for second test comes from here: Pandigital Numbers C# prime computation: using System; using System.Diagnostics; class Program { static int primes(int n) { uint i, j; int countprimes = 0; for (i = 1; i <= n; i++) { bool isprime = true; for (j = 2; j <= Math.Sqrt(i); j++) if ((i % j) == 0) { isprime = false; break; } if (isprime) countprimes++; } return countprimes; } static void Main(string[] args) { int n = int.Parse(Console.ReadLine()); Stopwatch sw = new Stopwatch(); sw.Start(); int res = primes(n); sw.Stop(); Console.WriteLine("I found {0} prime numbers between 0 and {1} in {2} msecs.", res, n, sw.ElapsedMilliseconds); Console.ReadKey(); } } C++ variant: #include <iostream> #include <ctime> int primes(unsigned long n) { unsigned long i, j; int countprimes = 0; for(i = 1; i <= n; i++) { int isprime = 1; for(j = 2; j < (i^(1/2)); j++) if(!(i%j)) { isprime = 0; break; } countprimes+= isprime; } return countprimes; } int main() { int n, res; cin>>n; unsigned int start = clock(); res = primes(n); int tprime = clock() - start; cout<<"\nI found "<<res<<" prime numbers between 1 and "<<n<<" in "<<tprime<<" msecs."; return 0; } When I ran the test trying to find primes < than 100,000, C# variant finished in 0.409 seconds and C++ variant in 5.553 seconds. When I ran them for 1,000,000 C# finished in 6.039 seconds and C++ in about 337 seconds. Pandigital test in C#: using System; using System.Diagnostics; class Program { static bool IsPandigital(int n) { int digits = 0; int count = 0; int tmp; for (; n > 0; n /= 10, ++count) { if ((tmp = digits) == (digits |= 1 << (n - ((n / 10) * 10) - 1))) return false; } return digits == (1 << count) - 1; } static void Main() { int pans = 0; Stopwatch sw = new Stopwatch(); sw.Start(); for (int i = 1; i <= 123456789; i++) { if (IsPandigital(i)) { pans++; } } sw.Stop(); Console.WriteLine("{0}pcs, {1}ms", pans, sw.ElapsedMilliseconds); Console.ReadKey(); } } Pandigital test in C++: #include <iostream> #include <ctime> using namespace std; int IsPandigital(int n) { int digits = 0; int count = 0; int tmp; for (; n > 0; n /= 10, ++count) { if ((tmp = digits) == (digits |= 1 << (n - ((n / 10) * 10) - 1))) return 0; } return digits == (1 << count) - 1; } int main() { int pans = 0; unsigned int start = clock(); for (int i = 1; i <= 123456789; i++) { if (IsPandigital(i)) { pans++; } } int ptime = clock() - start; cout<<"\nPans:"<<pans<<" time:"<<ptime; return 0; } C# variant runs in 29.906 seconds and C++ in about 36.298 seconds. I didn't touch any compiler switches and bot C# and C++ programs were compiled with debug options. Before I attempted to run the test I was worried that C# will lag well behind C++, but now it seems that there is a pretty big speed difference in C# favor. Can anybody explain this? C# is jitted and C++ is compiled native so it's normal that a C++ will be faster than a C# variant. Thanks for the answers!

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  • Elapsed time of running a C program

    - by yCalleecharan
    Hi, I would like to know what lines of C code to add to a program so that it tells me the total time that the program takes to run. I guess there should be counter initialization near the beginning of main and one after the main function ends. Is the right header clock.h? Thanks a lot... Update I have a Win Xp machine. Is it just adding clock() at the beginning and another clock() at the end of the program? Then I can estimate the time difference. Yes, you're right it's time.h. Here's my code: #include <stdio.h> #include <stdlib.h> #include <math.h> #include <share.h> #include <time.h> void f(long double fb[], long double fA, long double fB); int main() { clock_t start, end; start = clock(); const int ARRAY_SIZE = 11; long double* z = (long double*) malloc(sizeof (long double) * ARRAY_SIZE); int i; long double A, B; if (z == NULL) { printf("Out of memory\n"); exit(-1); } A = 0.5; B = 2; for (i = 0; i < ARRAY_SIZE; i++) { z[i] = 0; } z[1] = 5; f(z, A, B); for (i = 0; i < ARRAY_SIZE; i++) printf("z is %.16Le\n", z[i]); free(z); z = NULL; end = clock(); printf("Took %ld ticks\n", end-start); printf("Took %f seconds\n", (double)(end-start)/CLOCKS_PER_SEC); return 0; } void f(long double fb[], long double fA, long double fB) { fb[0] = fb[1]* fA; fb[1] = fb[1] - 1; return; } Some errors with MVS2008: testim.c(16) : error C2143: syntax error : missing ';' before 'const' testim.c(18) :error C2143: syntax error : missing ';' before 'type' testim.c(20) :error C2143: syntax error : missing ';' before 'type' testim.c(21) :error C2143: syntax error : missing ';' before 'type' testim.c(23) :error C2065: 'z' : undeclared identifier testim.c(23) :warning C4047: '==' : 'int' differs in levels of indirection from 'void *' testim.c(28) : error C2065: 'A' : undeclared identifier testim.c(28) : warning C4244: '=' : conversion from 'double' to 'int', possible loss of data and it goes to 28 errors. Note that I don't have any errors/warnings without your clock codes. LATEST NEWS: I unfortunately didn't get a good reply here. But after a search on Google, the code is working. Here it is: #include <stdio.h> #include <stdlib.h> #include <math.h> #include <share.h> #include <time.h> void f(long double fb[], long double fA, long double fB); int main() { clock_t start = clock(); const int ARRAY_SIZE = 11; long double* z = (long double*) malloc(sizeof (long double) * ARRAY_SIZE); int i; long double A, B; if (z == NULL) { printf("Out of memory\n"); exit(-1); } A = 0.5; B = 2; for (i = 0; i < ARRAY_SIZE; i++) { z[i] = 0; } z[1] = 5; f(z, A, B); for (i = 0; i < ARRAY_SIZE; i++) printf("z is %.16Le\n", z[i]); free(z); z = NULL; printf("Took %f seconds\n", ((double)clock()-start)/CLOCKS_PER_SEC); return 0; } void f(long double fb[], long double fA, long double fB) { fb[0] = fb[1]* fA; fb[1] = fb[1] - 1; return; } Cheers

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  • Compile error with initializer_list when trying to use it to initialize member value of class

    - by ilektron
    I am trying to make a class initializable from an initialization_list in a class constructor's constructor's initialization list. It works for a std::map, but not for my custom class. I don't see any difference other than templates are used in std::map. #include <iostream> #include <initializer_list> #include <string> #include <sstream> #include <map> using std::string; class text_thing { private: string m_text; public: text_thing() { } text_thing(text_thing& other); text_thing(std::initializer_list< std::pair<const string, const string> >& il); text_thing& operator=(std::initializer_list< std::pair<const string, const string> >& il); operator string() { return m_text; } }; class static_base { private: std::map<string, string> m_test_map; text_thing m_thing; static_base(); public: static static_base& getInstance() { static static_base instance; return instance; } string getText() { return (string)m_thing; } }; typedef std::pair<const string, const string> spair; text_thing::text_thing(text_thing& other) { m_text = other.m_text; } text_thing::text_thing(std::initializer_list< std::pair<const string, const string> >& il) { std::stringstream text_gen; for (auto& apair : il) { text_gen << "{" << apair.first << ", " << apair.second << "}" << std::endl; } } text_thing& text_thing::operator=(std::initializer_list< std::pair<const string, const string> >& il) { std::stringstream text_gen; for (auto& apair : il) { text_gen << "{" << apair.first << ", " << apair.second << "}" << std::endl; } return *this; } static_base::static_base() : m_test_map{{"test", "1"}, {"test2", "2"}}, // Compiler fine with this m_thing{{"test", "1"}, {"test2", "2"}} // Compiler doesn't like this { } int main() { std::cout << "Starting the program" << std::endl; std::cout << "The text thing: " << std::endl << static_base::getInstance().getText(); } I get this compiler output g++ -O0 -g3 -Wall -c -fmessage-length=0 -std=c++11 -MMD -MP -MF"static_base.d" -MT"static_base.d" -o "static_base.o" "../static_base.cpp" Finished building: ../static_base.cpp Building file: ../test.cpp Invoking: GCC C++ Compiler g++ -O0 -g3 -Wall -c -fmessage-length=0 -std=c++11 -MMD -MP -MF"test.d" -MT"test.d" -o "test.o" "../test.cpp" ../test.cpp: In constructor ‘static_base::static_base()’: ../test.cpp:94:40: error: no matching function for call to ‘text_thing::text_thing(<brace-enclosed initializer list>)’ m_thing{{"test", "1"}, {"test2", "2"}} ^ ../test.cpp:94:40: note: candidates are: ../test.cpp:72:1: note: text_thing::text_thing(std::initializer_list<std::pair<const std::basic_string<char>, const std::basic_string<char> > >&) text_thing::text_thing(std::initializer_list< std::pair<const string, const string> >& il) ^ ../test.cpp:72:1: note: candidate expects 1 argument, 2 provided ../test.cpp:67:1: note: text_thing::text_thing(text_thing&) text_thing::text_thing(text_thing& other) ^ ../test.cpp:67:1: note: candidate expects 1 argument, 2 provided ../test.cpp:23:2: note: text_thing::text_thing() text_thing() ^ ../test.cpp:23:2: note: candidate expects 0 arguments, 2 provided make: *** [test.o] Error 1 Output of gcc -v Using built-in specs. COLLECT_GCC=gcc COLLECT_LTO_WRAPPER=/usr/lib/gcc/x86_64-linux-gnu/4.8/lto-wrapper Target: x86_64-linux-gnu Configured with: ../src/configure -v --with-pkgversion='Ubuntu 4.8.1-2ubuntu1~13.04' --with-bugurl=file:///usr/share/doc/gcc-4.8/README.Bugs --enable-languages=c,c++,java,go,d,fortran,objc,obj-c++ --prefix=/usr --program-suffix=-4.8 --enable-shared --enable-linker-build-id --libexecdir=/usr/lib --without-included-gettext --enable-threads=posix --with-gxx-include-dir=/usr/include/c++/4.8 --libdir=/usr/lib --enable-nls --with-sysroot=/ --enable-clocale=gnu --enable-libstdcxx-debug --enable-libstdcxx-time=yes --enable-gnu-unique-object --enable-plugin --with-system-zlib --disable-browser-plugin --enable-java-awt=gtk --enable-gtk-cairo --with-java-home=/usr/lib/jvm/java-1.5.0-gcj-4.8-amd64/jre --enable-java-home --with-jvm-root-dir=/usr/lib/jvm/java-1.5.0-gcj-4.8-amd64 --with-jvm-jar-dir=/usr/lib/jvm-exports/java-1.5.0-gcj-4.8-amd64 --with-arch-directory=amd64 --with-ecj-jar=/usr/share/java/eclipse-ecj.jar --enable-objc-gc --enable-multiarch --disable-werror --with-arch-32=i686 --with-abi=m64 --with-multilib-list=m32,m64,mx32 --with-tune=generic --enable-checking=release --build=x86_64-linux-gnu --host=x86_64-linux-gnu --target=x86_64-linux-gnu Thread model: posix gcc version 4.8.1 (Ubuntu 4.8.1-2ubuntu1~13.04) It compiles fine with the std::map constructed this way, and if I modify the static_base to return the strings from the maps, all is fine and dandy. Please help me understand what is going on here.

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  • Application crashing when talking to oracle unless executable path contains spaces

    - by Lasse V. Karlsen
    We have an x-files problem with our .NET application. Or, rather, hybrid Win32 and .NET application. When it attempts to communicate with Oracle, it just dies. Vanishes. Goes to the big black void in the sky. No event log message, no exception, no nothing. If we simply ask the application to talk to a MS SQL Server instead, which has the effect of replacing the usage of OracleConnection and related classes with SqlConnection and related classes, it works as expected. Today we had a breakthrough. For some reason, a customer had figured out that by placing all the application files in a directory on his desktop, it worked as expected with Oracle as well. Moving the directory down to the root of the drive, or in C:\Temp or, well, around a bit, made the crash reappear. Basically it was 100% reproducable that the application worked if run from directory on desktop, and failed if run from directory in root. Today we figured out that the difference that counted was wether there was a space in the directory name or not. So, these directories would work: C:\Program Files\AppDir\Executable.exe C:\Temp Lemp\AppDir\Executable.exe C:\Documents and Settings\someuser\Desktop\AppDir\Executable.exe whereas these would not: C:\CompanyName\AppDir\Executable.exe C:\Programfiler\AppDir\Executable.exe <-- Program Files in norwegian C:\Temp\AppDir\Executable.exe I'm hoping someone reading this has seen similar behavior and have a "aha, you need to twiddle the frob on the oracle glitz driver configuration" or similar. Anyone? Followup #1: Ok, I've processed the procmon output now, both files from when I hit the button that attempts to open the window that triggers the cascade failure, and I've noticed that they keep track mostly, there's some smallish differences near the top of both files, and they they keep track a long way down. However, when one run fails, the other keeps going and the next few lines of the log output are these: ReadFile C:\oracle\product\10.2.0\db_1\BIN\orageneric10.dll SUCCESS Offset: 274 432, Length: 32 768, I/O Flags: Non-cached, Paging I/O, Synchronous Paging I/O ReadFile C:\oracle\product\10.2.0\db_1\BIN\orageneric10.dll SUCCESS Offset: 233 472, Length: 32 768, I/O Flags: Non-cached, Paging I/O, Synchronous Paging I/O After this, the working run continues to execute, and the other touches the mscorwks.dll files a few times before threads close down and the app closes. Thus, the failed run does not touch the above files. Followup #2: Figured I'd try to upgrade the oracle client drivers, but 10.2.0.1 is apparently the highest version available for Windows 2003 server and XP clients. Followup #3: Well, we've ended up with a black-box solution. Basically we found that the problem is somewhere related to XPO and Oracle. XPO has a system-table it manages, called XPObjectType, with three columns: Oid, TypeName and AssemblyName. Due to how Oracle is configured in the databases we talk to, the column names were OID, TYPENAME and ASSEMBLYNAME. This would ordinarily not be a problem, except that XPO talks to the schema information directly and checks if the table is there with the right column names, and XPO doesn't handle case differences so it sees a XPObjectType table with three unknown columns and none of those it expects. Exactly what XPO does now I don't really know, but if I dropped this table, and recreated it with the right case, using double quotes around all the column names to get the case right, the problem doesn't crop up. Exactly where the space in the folder name comes into this, I still have no idea, but this problem had two tiers: Stop the application from crashing at our customers, short-term solution Fix the bug, long-term solution Right now tier 1 is solved, tier 2 will be put back into the queue for now and prioritized. We're facing some bigger changes to our data tier anyway so this might not be a problem we need to solve, at least if all our Oracle-customers verify that the table-fix actually gets rid of the problem. I'll accept the answer by Dave Markle since though Process Monitor (the big brother of File Monitor) didn't actually pinpoint the problem, I was able to use it to determine that after my breakpoint in user-code where XPO had built up the query for this table, no I/O happened until all the entries for the application closing down was logged, which led me to believe it was this table that was the culprit, or at least influenced the problem somehow. If I manage to get to the real cause of this, I'll update the post.

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  • Find the set of largest contiguous rectangles to cover multiple areas

    - by joelpt
    I'm working on a tool called Quickfort for the game Dwarf Fortress. Quickfort turns spreadsheets in csv/xls format into a series of commands for Dwarf Fortress to carry out in order to plot a "blueprint" within the game. I am currently trying to optimally solve an area-plotting problem for the 2.0 release of this tool. Consider the following "blueprint" which defines plotting commands for a 2-dimensional grid. Each cell in the grid should either be dug out ("d"), channeled ("c"), or left unplotted ("."). Any number of distinct plotting commands might be present in actual usage. . d . d c c d d d d c c . d d d . c d d d d d c . d . d d c To minimize the number of instructions that need to be sent to Dwarf Fortress, I would like to find the set of largest contiguous rectangles that can be formed to completely cover, or "plot", all of the plottable cells. To be valid, all of a given rectangle's cells must contain the same command. This is a faster approach than Quickfort 1.0 took: plotting every cell individually as a 1x1 rectangle. This video shows the performance difference between the two versions. For the above blueprint, the solution looks like this: . 9 . 0 3 2 8 1 1 1 3 2 . 1 1 1 . 2 7 1 1 1 4 2 . 6 . 5 4 2 Each same-numbered rectangle above denotes a contiguous rectangle. The largest rectangles take precedence over smaller rectangles that could also be formed in their areas. The order of the numbering/rectangles is unimportant. My current approach is iterative. In each iteration, I build a list of the largest rectangles that could be formed from each of the grid's plottable cells by extending in all 4 directions from the cell. After sorting the list largest first, I begin with the largest rectangle found, mark its underlying cells as "plotted", and record the rectangle in a list. Before plotting each rectangle, its underlying cells are checked to ensure they are not yet plotted (overlapping a previous plot). We then start again, finding the largest remaining rectangles that can be formed and plotting them until all cells have been plotted as part of some rectangle. I consider this approach slightly more optimized than a dumb brute-force search, but I am wasting a lot of cycles (re)calculating cells' largest rectangles and checking underlying cells' states. Currently, this rectangle-discovery routine takes the lion's share of the total runtime of the tool, especially for large blueprints. I have sacrificed some accuracy for the sake of speed by only considering rectangles from cells which appear to form a rectangle's corner (determined using some neighboring-cell heuristics which aren't always correct). As a result of this 'optimization', my current code doesn't actually generate the above solution correctly, but it's close enough. More broadly, I consider the goal of largest-rectangles-first to be a "good enough" approach for this application. However I observe that if the goal is instead to find the minimum set (fewest number) of rectangles to completely cover multiple areas, the solution would look like this instead: . 3 . 5 6 8 1 3 4 5 6 8 . 3 4 5 . 8 2 3 4 5 7 8 . 3 . 5 7 8 This second goal actually represents a more optimal solution to the problem, as fewer rectangles usually means fewer commands sent to Dwarf Fortress. However, this approach strikes me as closer to NP-Hard, based on my limited math knowledge. Watch the video if you'd like to better understand the overall strategy; I have not addressed other aspects of Quickfort's process, such as finding the shortest cursor-path that plots all rectangles. Possibly there is a solution to this problem that coherently combines these multiple strategies. Help of any form would be appreciated.

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  • Which workaround to use for the following SQL deadlock?

    - by Marko
    I found a SQL deadlock scenario in my application during concurrency. I belive that the two statements that cause the deadlock are (note - I'm using LINQ2SQL and DataContext.ExecuteCommand(), that's where this.studioId.ToString() comes into play): exec sp_executesql N'INSERT INTO HQ.dbo.SynchronizingRows ([StudioId], [UpdatedRowId]) SELECT @p0, [t0].[Id] FROM [dbo].[UpdatedRows] AS [t0] WHERE NOT (EXISTS( SELECT NULL AS [EMPTY] FROM [dbo].[ReceivedUpdatedRows] AS [t1] WHERE ([t1].[StudioId] = @p0) AND ([t1].[UpdatedRowId] = [t0].[Id]) ))',N'@p0 uniqueidentifier',@p0='" + this.studioId.ToString() + "'; and exec sp_executesql N'INSERT INTO HQ.dbo.ReceivedUpdatedRows ([UpdatedRowId], [StudioId], [ReceiveDateTime]) SELECT [t0].[UpdatedRowId], @p0, GETDATE() FROM [dbo].[SynchronizingRows] AS [t0] WHERE ([t0].[StudioId] = @p0)',N'@p0 uniqueidentifier',@p0='" + this.studioId.ToString() + "'; The basic logic of my (client-server) application is this: Every time someone inserts or updates a row on the server side, I also insert a row into the table UpdatedRows, specifying the RowId of the modified row. When a client tries to synchronize data, it first copies all of the rows in the UpdatedRows table, that don't contain a reference row for the specific client in the table ReceivedUpdatedRows, to the table SynchronizingRows (the first statement taking part in the deadlock). Afterwards, during the synchronization I look for modified rows via lookup of the SynchronizingRows table. This step is required, otherwise if someone inserts new rows or modifies rows on the server side during synchronization I will miss them and won't get them during the next synchronization (explanation scenario to long to write here...). Once synchronization is complete, I insert rows to the ReceivedUpdatedRows table specifying that this client has received the UpdatedRows contained in the SynchronizingRows table (the second statement taking part in the deadlock). Finally I delete all rows from the SynchronizingRows table that belong to the current client. The way I see it, the deadlock is occuring on tables SynchronizingRows (abbreviation SR) and ReceivedUpdatedRows (abbreviation RUR) during steps 2 and 3 (one client is in step 2 and is inserting into SR and selecting from RUR; while another client is in step 3 inserting into RUR and selecting from SR). I googled a bit about SQL deadlocks and came to a conclusion that I have three options. Inorder to make a decision I need more input about each option/workaround: Workaround 1: The first advice given on the web about SQL deadlocks - restructure tables/queries so that deadlocks don't happen in the first place. Only problem with this is that with my IQ I don't see a way to do the synchronization logic any differently. If someone wishes to dwelve deeper into my current synchronization logic, how and why it is set up the way it is, I'll post a link for the explanation. Perhaps, with the help of someone smarter than me, it's possible to create a logic that is deadlock free. Workaround 2: The second most common advice seems to be the use of WITH(NOLOCK) hint. The problem with this is that NOLOCK might miss or duplicate some rows. Duplication is not a problem, but missing rows is catastrophic! Another option is the WITH(READPAST) hint. On the face of it, this seems to be a perfect solution. I really don't care about rows that other clients are inserting/modifying, because each row belongs only to a specific client, so I may very well skip locked rows. But the MSDN documentaion makes me a bit worried - "When READPAST is specified, both row-level and page-level locks are skipped". As I said, row-level locks would not be a problem, but page-level locks may very well be, since a page might contain rows that belong to multiple clients (including the current one). While there are lots of blog posts specifically mentioning that NOLOCK might miss rows, there seems to be none about READPAST (never) missing rows. This makes me skeptical and nervous to implement it, since there is no easy way to test it (implementing would be a piece of cake, just pop WITH(READPAST) into both statements SELECT clause and job done). Can someone confirm whether the READPAST hint can miss rows? Workaround 3: The final option is to use ALLOW_SNAPSHOT_ISOLATION and READ_COMMITED_SNAPSHOT. This would seem to be the only option to work 100% - at least I can't find any information that would contradict with it. But it is a little bit trickier to setup (I don't care much about the performance hit), because I'm using LINQ. Off the top of my head I probably need to manually open a SQL connection and pass it to the LINQ2SQL DataContext, etc... I haven't looked into the specifics very deeply. Mostly I would prefer option 2 if somone could only reassure me that READPAST will never miss rows concerning the current client (as I said before, each client has and only ever deals with it's own set of rows). Otherwise I'll likely have to implement option 3, since option 1 is probably impossible... I'll post the table definitions for the three tables as well, just in case: CREATE TABLE [dbo].[UpdatedRows]( [Id] [uniqueidentifier] NOT NULL ROWGUIDCOL DEFAULT NEWSEQUENTIALID() PRIMARY KEY CLUSTERED, [RowId] [uniqueidentifier] NOT NULL, [UpdateDateTime] [datetime] NOT NULL, ) ON [PRIMARY] GO CREATE NONCLUSTERED INDEX IX_RowId ON dbo.UpdatedRows ([RowId] ASC) WITH (STATISTICS_NORECOMPUTE = OFF, IGNORE_DUP_KEY = OFF, ALLOW_ROW_LOCKS = ON, ALLOW_PAGE_LOCKS = ON) ON [PRIMARY] GO CREATE TABLE [dbo].[ReceivedUpdatedRows]( [Id] [uniqueidentifier] NOT NULL ROWGUIDCOL DEFAULT NEWSEQUENTIALID() PRIMARY KEY NONCLUSTERED, [UpdatedRowId] [uniqueidentifier] NOT NULL REFERENCES [dbo].[UpdatedRows] ([Id]), [StudioId] [uniqueidentifier] NOT NULL REFERENCES, [ReceiveDateTime] [datetime] NOT NULL, ) ON [PRIMARY] GO CREATE CLUSTERED INDEX IX_Studios ON dbo.ReceivedUpdatedRows ([StudioId] ASC) WITH (STATISTICS_NORECOMPUTE = OFF, IGNORE_DUP_KEY = OFF, ALLOW_ROW_LOCKS = ON, ALLOW_PAGE_LOCKS = ON) ON [PRIMARY] GO CREATE TABLE [dbo].[SynchronizingRows]( [StudioId] [uniqueidentifier] NOT NULL [UpdatedRowId] [uniqueidentifier] NOT NULL REFERENCES [dbo].[UpdatedRows] ([Id]) PRIMARY KEY CLUSTERED ([StudioId], [UpdatedRowId]) ) ON [PRIMARY] GO PS! Studio = Client. PS2! I just noticed that the index definitions have ALLOW_PAGE_LOCK=ON. If I would turn it off, would that make any difference to READPAST? Are there any negative downsides for turning it off?

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  • Multiple windows services in a single project = mystery

    - by Remoh
    I'm having a bizarre issue that I haven't seen before and I'm thinking it MUST be something simple that I'm not seeing in my code. I have a project with 2 windows services defined. One I've called DataSyncService, the other SubscriptionService. Both are added to the same project installer. Both use a timer control from System.Timers. If I start both services together, they seem to work fine. The timers elapse at the appropriate time and everything looks okay. However, if I start either service individually, leaving the other stopped, everything goes haywire. The timer elapses constantly and on the wrong service. In other words, if I start the DataSyncService, the SubscriptionService timer elapses over and over. ...which is obviously strange. The setup is similar to what I've done in the past so I'm really stumped. I even tried deleting both service and starting over but it doesn't seem to make a difference. At this point, I'm thinking I've made a simple error in the way I'm defining the services and my brain just won't let me see it. It must be creating some sort of threading issue that causes one service to race when the other is stopped. Here the code.... From Program.cs: static void Main() { ServiceBase[] ServicesToRun; ServicesToRun = new ServiceBase[] { new DataSyncService(), new SubscriptionService() }; ServiceBase.Run(ServicesToRun); } From ProjectInstaller.designer.cs: private void InitializeComponent() { this.serviceProcessInstaller1 = new System.ServiceProcess.ServiceProcessInstaller(); this.dataSyncInstaller = new System.ServiceProcess.ServiceInstaller(); this.subscriptionInstaller = new System.ServiceProcess.ServiceInstaller(); // // serviceProcessInstaller1 // this.serviceProcessInstaller1.Account = System.ServiceProcess.ServiceAccount.LocalSystem; this.serviceProcessInstaller1.Password = null; this.serviceProcessInstaller1.Username = null; // // dataSyncInstaller // this.dataSyncInstaller.DisplayName = "Data Sync Service"; this.dataSyncInstaller.ServiceName = "DataSyncService"; this.dataSyncInstaller.StartType = System.ServiceProcess.ServiceStartMode.Automatic; // // subscriptionInstaller // this.subscriptionInstaller.DisplayName = "Subscription Service"; this.subscriptionInstaller.ServiceName = "SubscriptionService"; this.subscriptionInstaller.StartType = System.ServiceProcess.ServiceStartMode.Automatic; // // ProjectInstaller // this.Installers.AddRange(new System.Configuration.Install.Installer[] { this.serviceProcessInstaller1, this.dataSyncInstaller, this.subscriptionInstaller}); } private System.ServiceProcess.ServiceProcessInstaller serviceProcessInstaller1; private System.ServiceProcess.ServiceInstaller dataSyncInstaller; private System.ServiceProcess.ServiceInstaller subscriptionInstaller; From DataSyncService.cs: public static readonly int _defaultInterval = 43200000; //log4net.ILog log; public DataSyncService() { InitializeComponent(); //log = LogFactory.Instance.GetLogger(this); } protected override void OnStart(string[] args) { timer1.Interval = _defaultInterval; //GetInterval(); timer1.Enabled = true; EventLog.WriteEntry("MyProj", "Data Sync Service Started", EventLogEntryType.Information); //log.Info("Data Sync Service Started"); } private void timer1_Elapsed(object sender, System.Timers.ElapsedEventArgs e) { EventLog.WriteEntry("MyProj", "Data Sync Timer Elapsed.", EventLogEntryType.Information); } private void InitializeComponent() { this.timer1 = new System.Timers.Timer(); ((System.ComponentModel.ISupportInitialize)(this.timer1)).BeginInit(); // // timer1 // this.timer1.Enabled = true; this.timer1.Elapsed += new System.Timers.ElapsedEventHandler(this.timer1_Elapsed); // // DataSyncService // this.ServiceName = "DataSyncService"; ((System.ComponentModel.ISupportInitialize)(this.timer1)).EndInit(); } From SubscriptionService: public static readonly int _defaultInterval = 300000; //log4net.ILog log; public SubscriptionService() { InitializeComponent(); } protected override void OnStart(string[] args) { timer1.Interval = _defaultInterval; //GetInterval(); timer1.Enabled = true; EventLog.WriteEntry("MyProj", "Subscription Service Started", EventLogEntryType.Information); //log.Info("Subscription Service Started"); } private void timer1_Elapsed(object sender, System.Timers.ElapsedEventArgs e) { EventLog.WriteEntry("MyProj", "Subscription Service Time Elapsed", EventLogEntryType.Information); } private void InitializeComponent() //in designer { this.timer1 = new System.Timers.Timer(); ((System.ComponentModel.ISupportInitialize)(this.timer1)).BeginInit(); // // timer1 // this.timer1.Enabled = true; this.timer1.Elapsed += new System.Timers.ElapsedEventHandler(this.timer1_Elapsed); // // SubscriptionService // this.ServiceName = "SubscriptionService"; ((System.ComponentModel.ISupportInitialize)(this.timer1)).EndInit(); } Again, the problem is that the timer1_elapsed handler runs constantly when only one of the services is started. And it's the handler on the OPPOSITE service. Anybody see anything?

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  • Re-opening closed dialog puts it in the start position after moving it

    - by semmelbroesel
    I'm using multiple instances of the JQuery-UI dialog along with draggable and resizable. Whenever I drag or resize one of the dialog boxes, the position and dimensions of the current box are saved to a database and then loaded the next time the page opens. This is working well. However, when I close a dialog box and re-open it using a button, jQuery sets the position of the box back to its original location in the center of the screen. Furthermore, I use a show effect to slide the box off to the left on closing and in from the left on re-opening. I found two ways to update its position when the slide in animation is done, however, it still slides into the center of the screen, and I have yet to find a way to get it to slide in towards the location it is supposed to have. Here are the parts of the code that play a part in this: $('.box').dialog({ closeOnEscape: false, hide: { effect: "drop", direction: 'left' }, beforeClose: function (evt, ui){ var $this = $(this); SavePos($this); // saves the dimensions to db }, dragStop: function() { var $this = $(this); SavePos($this); }, resizeStop: function() { var $this = $(this); SavePos($this); }, open: function() { var $this = $(this); $this.dialog('option', { show: { effect: "drop", direction: 'left'} } ); if (init) // meaning only load this code when the page has finished initializing { // tried it both ways - set the position before and after the effect - no success UpdatePos($this.attr('id')); // I tried this section with promise() and effect / complete - I found no difference $this.parent().promise().done(function() { UpdatePos($this.attr('id')); }); } } }); function UpdatePos(key) { // boxpos is an object holding the position for each box by the box's id var $this = $('#' + key); //console.log('updating pos for box ' + boxid); if ($this && $this.hasClass('ui-dialog-content')) { //console.log($this.dialog('widget').css('left')); $this.dialog('option', { width: boxpos[key].width, height: boxpos[key].height }); $this.dialog('widget').css({ left: boxpos[key].left + 'px', top: boxpos[key].top + 'px' }); //console.log('finished updating pos'); //console.log($this.dialog('widget').css('left')); } } The button that re-opens the box has this code on it to make that happen: var $box = $('#boxid'); if ($box) { if ($box.dialog('isOpen')) { $box.dialog('moveToTop'); } else { $box.dialog("open") } } I don't know what jQuery-UI does to the box as it hides it (other than display:none) or to make it slide in, so maybe there's something I'm missing here that might help... Basically, I need JQuery to remember the box' position and put the box back into that location when it is re-opened. It took me days to make it this far, but this is one obstacle I have yet to overcome. Maybe there's a different way I can re-open the box? Thanks! EDIT: Forgot - this issue ONLY happens when I use my UpdatePos function to set the location of a box (i.e. on page load). When I drag a box with my mouse, close it, and re-open it, everything works. So I'm guessing there's one more storage location for the box' position that I'm missing here... EDIT2: After more messing with it, my code for debugging now looks like this: open: function() { var $this = $(this); console.log('box open'); console.log($this.dialog('widget').position()); // { top=0, left=-78.5} console.log($this.dialog('widget').css('left')); $this.dialog('option', { show: { effect: "drop", direction: 'left'} } ); if (init) { UpdatePos($this.attr('id')); $this.parent().promise().done(function() { console.log($this.dialog('widget').position()); // { top=313, left=641.5} console.log($this.dialog('widget').css('left')); UpdatePos($this.attr('id')); console.log($this.dialog('widget').position()); // { top=121, left=107} console.log($this.dialog('widget').css('left')); }); } The results I'm getting are: box open Object { top=0, left=-78.5} -78.5px Object { top=313, left=641.5} 641.5px Object { top=121, left=107} 107px So looks to me as if the widget is being moved off screen (left=-78.5) and then moved for the animation, and then my code moves it into the location that it should be in (121/107). The position() results for $box.position() or $box.dialog().position() do not change during this debugging section. Maybe this will help someone here - I'm still out of ideas here... ... and I just discovered that when I drag the item around myself, then close and re-open it, it is very unpredictable. Sometimes, it will end up in the correct location horizontally, but not vertically. Sometimes, it will end up back in the center of the screen...

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  • Rails: Problem with routes and special Action.

    - by Newbie
    Hello! Sorry for this question but I can't find my error! In my Project I have my model called "team". A User can create a "team" or a "contest". The difference between this both is, that contest requires more data than a normal team. So I created the columns in my team table. Well... I also created a new view called create_contest.html.erb : <h1>New team content</h1> <% form_for @team, :url => { :action => 'create_content' } do |f| %> <%= f.error_messages %> <p> <%= f.label :name %><br /> <%= f.text_field :name %> </p> <p> <%= f.label :description %><br /> <%= f.text_area :description %> </p> <p> <%= f.label :url %><br /> <%= f.text_fiels :url %> </p> <p> <%= f.label :contact_name %><br /> <%= f.text_fiels :contact_name %> </p> <p> <%= f.submit 'Create' %> </p> <% end %> In my teams_controller, I created following functions: def new_contest end def create_contest if @can_create @team = Team.new(params[:team]) @team.user_id = current_user.id respond_to do |format| if @team.save format.html { redirect_to(@team, :notice => 'Contest was successfully created.') } format.xml { render :xml => @team, :status => :created, :location => @team } else format.html { render :action => "new" } format.xml { render :xml => @team.errors, :status => :unprocessable_entity } end end else redirect_back_or_default('/') end end Now, I want on my teams/new.html.erb a link to "new_contest.html.erb". So I did: <%= link_to 'click here for new contest!', new_contest_team_path %> When I go to the /teams/new.html.erb page, I get following error: undefined local variable or method `new_contest_team_path' for #<ActionView::Base:0x16fc4f7> So I changed in my routes.rb, map.resources :teams to map.resources :teams, :member=>{:new_contest => :get} Now I get following error: new_contest_team_url failed to generate from {:controller=>"teams", :action=>"new_contest"} - you may have ambiguous routes, or you may need to supply additional parameters for this route. content_url has the following required parameters: ["teams", :id, "new_contest"] - are they all satisfied? I don't think adding :member => {...} is the right way doing this. So, can you tell me what to do? I want to have an URL like /teams/new-contest or something. My next question: what to do (after fixing the first problem), to validate presentence of all fields for new_contest.html.erb? In my normal new.html.erb, a user does not need all the data. But in new_contest.html.erb he does. Is there a way to make a validates_presence_of only for one action (in this case new_contest)? UPDATE: Now, I removed my :member part from my routes.rb and wrote: map.new_contest '/teams/contest/new', :controller => 'teams', :action => 'new_contest' Now, clicking on my link, it redirects me to /teams/contest/new - like I wanted - but I get another error called: Called id for nil, which would mistakenly be 4 -- if you really wanted the id of nil, use object_id I think this error is cause of @team at <% form_for @team, :url => { :action => 'create_content_team' } do |f| %> What to do for solving this error?

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  • Objects in Java ArrayList don't get updated.

    - by Sbm007
    This is going to be a very long post, hopefully you can understand what I'm talking about and I appreciate any help. Thanks Basically, I've created a personal, non-commercial project (which I don't plan to release) that can read ZIP and RAR files. It can only read the contents in the archive, the folders inside, the files inside the folders and its properties (such as last modified date, last modified time, CRC checksum, uncompressed size, compressed size and file name). It can't extract files either, so it's really a ZIP/RAR viewer if you may. Anyway that's slightly irrelevant to my problem but I thought I'd give you some background info. Now for my problem: I can successfully list all the folders and files inside a ZIP archive, so now I want to take that raw input and link it together in some useful way. I made 2 classes: ArchiveFile (represents a file inside a ZIP) and ArchiveFolder (represents a folder inside a ZIP). They both have some useful methods such as getLastModifiedDate, getName, getPath and so on. But the difference is that ArchiveFolder can hold an ArrayList of ArchiveFile's and additional ArchiveFolder's (think of this as files and folders inside a folder). Now I want to populate my raw input into one root ArchiveFolder, which will have all the files in the root dir of the ZIP in the ArchiveFile's ArrayList and any additional folders in the root dir of the ZIP in the ArchiveFolder's ArrayList (and this process can continue on like this like a chain reaction (more files/folders in that ArchiveFolder etc etc). So I came up with the following code: while (archive.hasMore()) { String path = ""; ArchiveFolder current = root; String[] contents = archive.getName().split("/"); for (int x = 0; x < contents.length; ++x) { if (x == (contents.length - 1) && !archive.getName().endsWith("/")) { // If on last item and item is a file path += contents[x]; // Update final path ArchiveFile file = new ArchiveFile(path, contents[x], archive.getUncompressedSize(), archive.getCompressedSize(), archive.getModifiedTime(), archive.getModifiedDate(), archive.getCRC()); current.addFile(file); // Create and add the file to the current ArchiveFolder } else if (x == (contents.length - 1)) { // Else if we are on last item and it is a folder path += contents[x] + "/"; // Update final path ArchiveFolder folder = new ArchiveFolder(path, contents[x], archive.getModifiedTime(), archive.getModifiedDate()); current.addFolder(folder); // Create and add this folder to the current ArchiveFile } else { // Else if we are still traversing through the path path += contents[x] + "/"; // Update path ArchiveFolder folder = new ArchiveFolder(path, contents[x]); current.addFolder(folder); // Create and add folder (remember we do not know the modified date/time as all we know is the path, so we can deduce the name only) current = folder; // Update current ArchiveFolder to the newly created one for the next iteration of the for loop } } archive.getNext(); } Assume that root is the root ArchiveFolder (initially empty). And that archive.getName() returns the name of the current file OR folder in the following fashion: file.txt or folder1/file2.txt or folder4/folder2/ (this is a empty folder) etc. So basically the relative path from the root of the ZIP archive. Please read through the comments in the above code to familiarize yourself with it. Also assume that the addFolder method in an ArchiveFile, only adds the folder if it doesn't exist already (so there are no multiple folders) and it also updates the time and date of an existing folder if it is blank (ie it was a intermediate folder we only knew the name of, but now we know its details). The code for addFolder is (pretty self-explanitory): public void addFolder(ArchiveFolder folder) { int loc = folders.indexOf(folder); // folders is the ArrayList containing ArchiveFolder's if (loc == -1) { folders.add(folder); } else { ArchiveFolder real = folders.get(loc); if (real.time == null) { real.setTime(folder.getTime()); real.setDate(folder.getDate()); } } } So I can't see anything wrong with the code, it works and after finishing, the root ArchiveFolder contains all the files in the root of the ZIP as I want it to, and it contains all the direcories in the root folder as I want it to. So you'd think it works as expected, but no the ArchiveFolder's in the root folder don't contain the data inside those 'child' folders, it's just a blank folder with no additional files and folders (while it does really contain some more files/folders when viewed in WinZip). After debugging using Eclipse, the for loop does iterate through all the files (even those not included above), so this led me to believe that there is a problem with this line of the code: current = folder; What it does is, it updates the current folder (used as an intermediate by the loop) to the newly added folder. I thought Java passed by reference and thus all new operations and new additions in future ArchiveFile's and ArchiveFolder's are automatically updated, and parent ArchiveFolder's will be updated accordingly. But that does not appear to be the case? I know this is a long ass post and I really hope anyone can help me out with this. Thanks in advance.

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  • Accessing local variable doesn't improve performance

    - by NicMagnier
    The short version Why is this code: var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; More performant than this one? var index = Math.floor(ref_index) * 4; The long version This week, the author of Impact js published an article about some rendering issue: http://www.phoboslab.org/log/2012/09/drawing-pixels-is-hard In the article there was the source of a function to scale an image by accessing pixels in the canvas. I wanted to suggest some traditional ways to optimize this kind of code so that the scaling would be shorter at loading time. But after testing it my result was most of the time worst that the original function. Guessing this was the JavaScript engine that was doing some smart optimization I tried to understand a bit more what was going on so I did a bunch of test. But my results are quite confusing and I would need some help to understand what's going on. I have a test page here: http://www.mx981.com/stuff/resize_bench/test.html jsPerf: http://jsperf.com/local-variable-due-to-the-scope-lookup To start the test, click the picture and the results will appear in the console. There are three different versions: The original code: for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; var indexScaled = (y * widthScaled + x) * 4; scaledPixels.data[ indexScaled ] = origPixels.data[ index ]; scaledPixels.data[ indexScaled+1 ] = origPixels.data[ index+1 ]; scaledPixels.data[ indexScaled+2 ] = origPixels.data[ index+2 ]; scaledPixels.data[ indexScaled+3 ] = origPixels.data[ index+3 ]; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance One of my attempt to optimize it: var ref_index = 0; var ref_indexScaled = 0 var ref_step = 1 / scale; for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = Math.floor(ref_index) * 4; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+1 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+2 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+3 ]; ref_index+= ref_step; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance The same optimized code but with recalculating the index variable each time (Hybrid) var ref_index = 0; var ref_indexScaled = 0 var ref_step = 1 / scale; for( var y = 0; y < heightScaled; y++ ) { for( var x = 0; x < widthScaled; x++ ) { var index = (Math.floor(y / scale) * img.width + Math.floor(x / scale)) * 4; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+1 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+2 ]; scaledPixels.data[ ref_indexScaled++ ] = origPixels.data[ index+3 ]; ref_index+= ref_step; } } jsPerf: http://jsperf.com/so-accessing-local-variable-doesn-t-improve-performance The only difference in the two last one is the calculation of the 'index' variable. And to my surprise the optimized version is slower in most browsers (except opera). Results of personal testing (not the jsPerf tests): Opera Original: 8668ms Optimized: 932ms Hybrid: 8696ms Chrome Original: 139ms Optimized: 145ms Hybrid: 136ms Safari Original: 433ms Optimized: 853ms Hybrid: 451ms Firefox Original: 343ms Optimized: 422ms Hybrid: 350ms After digging around, it seems an usual good practice is to access mainly local variable due to the scope lookup. Because The optimized version only call one local variable it should be faster that the Hybrid code which call multiple variable and object in addition to the various operation involved. So why the "optimized" version is slower? I thought that it might be because some JavaScript engine don't optimize the Optimized version because it is not hot enough but after using --trace-opt in chrome, it seems all version are properly compiled by V8. At this point I am a bit clueless and wonder if somebody would know what is going on? I did also some more test cases in this page: http://www.mx981.com/stuff/resize_bench/index.html

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  • Integrating JavaScript Unit Tests with Visual Studio

    - by Stephen Walther
    Modern ASP.NET web applications take full advantage of client-side JavaScript to provide better interactivity and responsiveness. If you are building an ASP.NET application in the right way, you quickly end up with lots and lots of JavaScript code. When writing server code, you should be writing unit tests. One big advantage of unit tests is that they provide you with a safety net that enable you to safely modify your existing code – for example, fix bugs, add new features, and make performance enhancements -- without breaking your existing code. Every time you modify your code, you can execute your unit tests to verify that you have not broken anything. For the same reason that you should write unit tests for your server code, you should write unit tests for your client code. JavaScript is just as susceptible to bugs as C#. There is no shortage of unit testing frameworks for JavaScript. Each of the major JavaScript libraries has its own unit testing framework. For example, jQuery has QUnit, Prototype has UnitTestJS, YUI has YUI Test, and Dojo has Dojo Objective Harness (DOH). The challenge is integrating a JavaScript unit testing framework with Visual Studio. Visual Studio and Visual Studio ALM provide fantastic support for server-side unit tests. You can easily view the results of running your unit tests in the Visual Studio Test Results window. You can set up a check-in policy which requires that all unit tests pass before your source code can be committed to the source code repository. In addition, you can set up Team Build to execute your unit tests automatically. Unfortunately, Visual Studio does not provide “out-of-the-box” support for JavaScript unit tests. MS Test, the unit testing framework included in Visual Studio, does not support JavaScript unit tests. As soon as you leave the server world, you are left on your own. The goal of this blog entry is to describe one approach to integrating JavaScript unit tests with MS Test so that you can execute your JavaScript unit tests side-by-side with your C# unit tests. The goal is to enable you to execute JavaScript unit tests in exactly the same way as server-side unit tests. You can download the source code described by this project by scrolling to the end of this blog entry. Rejected Approach: Browser Launchers One popular approach to executing JavaScript unit tests is to use a browser as a test-driver. When you use a browser as a test-driver, you open up a browser window to execute and view the results of executing your JavaScript unit tests. For example, QUnit – the unit testing framework for jQuery – takes this approach. The following HTML page illustrates how you can use QUnit to create a unit test for a function named addNumbers(). <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title>Using QUnit</title> <link rel="stylesheet" href="http://github.com/jquery/qunit/raw/master/qunit/qunit.css" type="text/css" /> </head> <body> <h1 id="qunit-header">QUnit example</h1> <h2 id="qunit-banner"></h2> <div id="qunit-testrunner-toolbar"></div> <h2 id="qunit-userAgent"></h2> <ol id="qunit-tests"></ol> <div id="qunit-fixture">test markup, will be hidden</div> <script type="text/javascript" src="http://code.jquery.com/jquery-latest.js"></script> <script type="text/javascript" src="http://github.com/jquery/qunit/raw/master/qunit/qunit.js"></script> <script type="text/javascript"> // The function to test function addNumbers(a, b) { return a+b; } // The unit test test("Test of addNumbers", function () { equals(4, addNumbers(1,3), "1+3 should be 4"); }); </script> </body> </html> This test verifies that calling addNumbers(1,3) returns the expected value 4. When you open this page in a browser, you can see that this test does, in fact, pass. The idea is that you can quickly refresh this QUnit HTML JavaScript test driver page in your browser whenever you modify your JavaScript code. In other words, you can keep a browser window open and keep refreshing it over and over while you are developing your application. That way, you can know very quickly whenever you have broken your JavaScript code. While easy to setup, there are several big disadvantages to this approach to executing JavaScript unit tests: You must view your JavaScript unit test results in a different location than your server unit test results. The JavaScript unit test results appear in the browser and the server unit test results appear in the Visual Studio Test Results window. Because all of your unit test results don’t appear in a single location, you are more likely to introduce bugs into your code without noticing it. Because your unit tests are not integrated with Visual Studio – in particular, MS Test -- you cannot easily include your JavaScript unit tests when setting up check-in policies or when performing automated builds with Team Build. A more sophisticated approach to using a browser as a test-driver is to automate the web browser. Instead of launching the browser and loading the test code yourself, you use a framework to automate this process. There are several different testing frameworks that support this approach: · Selenium – Selenium is a very powerful framework for automating browser tests. You can create your tests by recording a Firefox session or by writing the test driver code in server code such as C#. You can learn more about Selenium at http://seleniumhq.org/. LTAF – The ASP.NET team uses the Lightweight Test Automation Framework to test JavaScript code in the ASP.NET framework. You can learn more about LTAF by visiting the project home at CodePlex: http://aspnet.codeplex.com/releases/view/35501 jsTestDriver – This framework uses Java to automate the browser. jsTestDriver creates a server which can be used to automate multiple browsers simultaneously. This project is located at http://code.google.com/p/js-test-driver/ TestSwam – This framework, created by John Resig, uses PHP to automate the browser. Like jsTestDriver, the framework creates a test server. You can open multiple browsers that are automated by the test server. Learn more about TestSwarm by visiting the following address: https://github.com/jeresig/testswarm/wiki Yeti – This is the framework introduced by Yahoo for automating browser tests. Yeti uses server-side JavaScript and depends on Node.js. Learn more about Yeti at http://www.yuiblog.com/blog/2010/08/25/introducing-yeti-the-yui-easy-testing-interface/ All of these frameworks are great for integration tests – however, they are not the best frameworks to use for unit tests. In one way or another, all of these frameworks depend on executing tests within the context of a “living and breathing” browser. If you create an ASP.NET Unit Test then Visual Studio will launch a web server before executing the unit test. Why is launching a web server so bad? It is not the worst thing in the world. However, it does introduce dependencies that prevent your code from being tested in isolation. One of the defining features of a unit test -- versus an integration test – is that a unit test tests code in isolation. Another problem with launching a web server when performing unit tests is that launching a web server can be slow. If you cannot execute your unit tests quickly, you are less likely to execute your unit tests each and every time you make a code change. You are much more likely to fall into the pit of failure. Launching a browser when performing a JavaScript unit test has all of the same disadvantages as launching a web server when performing an ASP.NET unit test. Instead of testing a unit of JavaScript code in isolation, you are testing JavaScript code within the context of a particular browser. Using the frameworks listed above for integration tests makes perfect sense. However, I want to consider a different approach for creating unit tests for JavaScript code. Using Server-Side JavaScript for JavaScript Unit Tests A completely different approach to executing JavaScript unit tests is to perform the tests outside of any browser. If you really want to test JavaScript then you should test JavaScript and leave the browser out of the testing process. There are several ways that you can execute JavaScript on the server outside the context of any browser: Rhino – Rhino is an implementation of JavaScript written in Java. The Rhino project is maintained by the Mozilla project. Learn more about Rhino at http://www.mozilla.org/rhino/ V8 – V8 is the open-source Google JavaScript engine written in C++. This is the JavaScript engine used by the Chrome web browser. You can download V8 and embed it in your project by visiting http://code.google.com/p/v8/ JScript – JScript is the JavaScript Script Engine used by Internet Explorer (up to but not including Internet Explorer 9), Windows Script Host, and Active Server Pages. Internet Explorer is still the most popular web browser. Therefore, I decided to focus on using the JScript Script Engine to execute JavaScript unit tests. Using the Microsoft Script Control There are two basic ways that you can pass JavaScript to the JScript Script Engine and execute the code: use the Microsoft Windows Script Interfaces or use the Microsoft Script Control. The difficult and proper way to execute JavaScript using the JScript Script Engine is to use the Microsoft Windows Script Interfaces. You can learn more about the Script Interfaces by visiting http://msdn.microsoft.com/en-us/library/t9d4xf28(VS.85).aspx The main disadvantage of using the Script Interfaces is that they are difficult to use from .NET. There is a great series of articles on using the Script Interfaces from C# located at http://www.drdobbs.com/184406028. I picked the easier alternative and used the Microsoft Script Control. The Microsoft Script Control is an ActiveX control that provides a higher level abstraction over the Window Script Interfaces. You can download the Microsoft Script Control from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac After you download the Microsoft Script Control, you need to add a reference to it to your project. Select the Visual Studio menu option Project, Add Reference to open the Add Reference dialog. Select the COM tab and add the Microsoft Script Control 1.0. Using the Script Control is easy. You call the Script Control AddCode() method to add JavaScript code to the Script Engine. Next, you call the Script Control Run() method to run a particular JavaScript function. The reference documentation for the Microsoft Script Control is located at the MSDN website: http://msdn.microsoft.com/en-us/library/aa227633%28v=vs.60%29.aspx Creating the JavaScript Code to Test To keep things simple, let’s imagine that you want to test the following JavaScript function named addNumbers() which simply adds two numbers together: MvcApplication1\Scripts\Math.js function addNumbers(a, b) { return 5; } Notice that the addNumbers() method always returns the value 5. Right-now, it will not pass a good unit test. Create this file and save it in your project with the name Math.js in your MVC project’s Scripts folder (Save the file in your actual MVC application and not your MVC test application). Creating the JavaScript Test Helper Class To make it easier to use the Microsoft Script Control in unit tests, we can create a helper class. This class contains two methods: LoadFile() – Loads a JavaScript file. Use this method to load the JavaScript file being tested or the JavaScript file containing the unit tests. ExecuteTest() – Executes the JavaScript code. Use this method to execute a JavaScript unit test. Here’s the code for the JavaScriptTestHelper class: JavaScriptTestHelper.cs   using System; using System.IO; using Microsoft.VisualStudio.TestTools.UnitTesting; using MSScriptControl; namespace MvcApplication1.Tests { public class JavaScriptTestHelper : IDisposable { private ScriptControl _sc; private TestContext _context; /// <summary> /// You need to use this helper with Unit Tests and not /// Basic Unit Tests because you need a Test Context /// </summary> /// <param name="testContext">Unit Test Test Context</param> public JavaScriptTestHelper(TestContext testContext) { if (testContext == null) { throw new ArgumentNullException("TestContext"); } _context = testContext; _sc = new ScriptControl(); _sc.Language = "JScript"; _sc.AllowUI = false; } /// <summary> /// Load the contents of a JavaScript file into the /// Script Engine. /// </summary> /// <param name="path">Path to JavaScript file</param> public void LoadFile(string path) { var fileContents = File.ReadAllText(path); _sc.AddCode(fileContents); } /// <summary> /// Pass the path of the test that you want to execute. /// </summary> /// <param name="testMethodName">JavaScript function name</param> public void ExecuteTest(string testMethodName) { dynamic result = null; try { result = _sc.Run(testMethodName, new object[] { }); } catch { var error = ((IScriptControl)_sc).Error; if (error != null) { var description = error.Description; var line = error.Line; var column = error.Column; var text = error.Text; var source = error.Source; if (_context != null) { var details = String.Format("{0} \r\nLine: {1} Column: {2}", source, line, column); _context.WriteLine(details); } } throw new AssertFailedException(error.Description); } } public void Dispose() { _sc = null; } } }     Notice that the JavaScriptTestHelper class requires a Test Context to be instantiated. For this reason, you can use the JavaScriptTestHelper only with a Visual Studio Unit Test and not a Basic Unit Test (These are two different types of Visual Studio project items). Add the JavaScriptTestHelper file to your MVC test application (for example, MvcApplication1.Tests). Creating the JavaScript Unit Test Next, we need to create the JavaScript unit test function that we will use to test the addNumbers() function. Create a folder in your MVC test project named JavaScriptTests and add the following JavaScript file to this folder: MvcApplication1.Tests\JavaScriptTests\MathTest.js /// <reference path="JavaScriptUnitTestFramework.js"/> function testAddNumbers() { // Act var result = addNumbers(1, 3); // Assert assert.areEqual(4, result, "addNumbers did not return right value!"); }   The testAddNumbers() function takes advantage of another JavaScript library named JavaScriptUnitTestFramework.js. This library contains all of the code necessary to make assertions. Add the following JavaScriptnitTestFramework.js to the same folder as the MathTest.js file: MvcApplication1.Tests\JavaScriptTests\JavaScriptUnitTestFramework.js var assert = { areEqual: function (expected, actual, message) { if (expected !== actual) { throw new Error("Expected value " + expected + " is not equal to " + actual + ". " + message); } } }; There is only one type of assertion supported by this file: the areEqual() assertion. Most likely, you would want to add additional types of assertions to this file to make it easier to write your JavaScript unit tests. Deploying the JavaScript Test Files This step is non-intuitive. When you use Visual Studio to run unit tests, Visual Studio creates a new folder and executes a copy of the files in your project. After you run your unit tests, your Visual Studio Solution will contain a new folder named TestResults that includes a subfolder for each test run. You need to configure Visual Studio to deploy your JavaScript files to the test run folder or Visual Studio won’t be able to find your JavaScript files when you execute your unit tests. You will get an error that looks something like this when you attempt to execute your unit tests: You can configure Visual Studio to deploy your JavaScript files by adding a Test Settings file to your Visual Studio Solution. It is important to understand that you need to add this file to your Visual Studio Solution and not a particular Visual Studio project. Right-click your Solution in the Solution Explorer window and select the menu option Add, New Item. Select the Test Settings item and click the Add button. After you create a Test Settings file for your solution, you can indicate that you want a particular folder to be deployed whenever you perform a test run. Select the menu option Test, Edit Test Settings to edit your test configuration file. Select the Deployment tab and select your MVC test project’s JavaScriptTest folder to deploy. Click the Apply button and the Close button to save the changes and close the dialog. Creating the Visual Studio Unit Test The very last step is to create the Visual Studio unit test (the MS Test unit test). Add a new unit test to your MVC test project by selecting the menu option Add New Item and selecting the Unit Test project item (Do not select the Basic Unit Test project item): The difference between a Basic Unit Test and a Unit Test is that a Unit Test includes a Test Context. We need this Test Context to use the JavaScriptTestHelper class that we created earlier. Enter the following test method for the new unit test: [TestMethod] public void TestAddNumbers() { var jsHelper = new JavaScriptTestHelper(this.TestContext); // Load JavaScript files jsHelper.LoadFile("JavaScriptUnitTestFramework.js"); jsHelper.LoadFile(@"..\..\..\MvcApplication1\Scripts\Math.js"); jsHelper.LoadFile("MathTest.js"); // Execute JavaScript Test jsHelper.ExecuteTest("testAddNumbers"); } This code uses the JavaScriptTestHelper to load three files: JavaScripUnitTestFramework.js – Contains the assert functions. Math.js – Contains the addNumbers() function from your MVC application which is being tested. MathTest.js – Contains the JavaScript unit test function. Next, the test method calls the JavaScriptTestHelper ExecuteTest() method to execute the testAddNumbers() JavaScript function. Running the Visual Studio JavaScript Unit Test After you complete all of the steps described above, you can execute the JavaScript unit test just like any other unit test. You can use the keyboard combination CTRL-R, CTRL-A to run all of the tests in the current Visual Studio Solution. Alternatively, you can use the buttons in the Visual Studio toolbar to run the tests: (Unfortunately, the Run All Impacted Tests button won’t work correctly because Visual Studio won’t detect that your JavaScript code has changed. Therefore, you should use either the Run Tests in Current Context or Run All Tests in Solution options instead.) The results of running the JavaScript tests appear side-by-side with the results of running the server tests in the Test Results window. For example, if you Run All Tests in Solution then you will get the following results: Notice that the TestAddNumbers() JavaScript test has failed. That is good because our addNumbers() function is hard-coded to always return the value 5. If you double-click the failing JavaScript test, you can view additional details such as the JavaScript error message and the line number of the JavaScript code that failed: Summary The goal of this blog entry was to explain an approach to creating JavaScript unit tests that can be easily integrated with Visual Studio and Visual Studio ALM. I described how you can use the Microsoft Script Control to execute JavaScript on the server. By taking advantage of the Microsoft Script Control, we were able to execute our JavaScript unit tests side-by-side with all of our other unit tests and view the results in the standard Visual Studio Test Results window. You can download the code discussed in this blog entry from here: http://StephenWalther.com/downloads/Blog/JavaScriptUnitTesting/JavaScriptUnitTests.zip Before running this code, you need to first install the Microsoft Script Control which you can download from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac

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  • Top things web developers should know about the Visual Studio 2013 release

    - by Jon Galloway
    ASP.NET and Web Tools for Visual Studio 2013 Release NotesASP.NET and Web Tools for Visual Studio 2013 Release NotesSummary for lazy readers: Visual Studio 2013 is now available for download on the Visual Studio site and on MSDN subscriber downloads) Visual Studio 2013 installs side by side with Visual Studio 2012 and supports round-tripping between Visual Studio versions, so you can try it out without committing to a switch Visual Studio 2013 ships with the new version of ASP.NET, which includes ASP.NET MVC 5, ASP.NET Web API 2, Razor 3, Entity Framework 6 and SignalR 2.0 The new releases ASP.NET focuses on One ASP.NET, so core features and web tools work the same across the platform (e.g. adding ASP.NET MVC controllers to a Web Forms application) New core features include new templates based on Bootstrap, a new scaffolding system, and a new identity system Visual Studio 2013 is an incredible editor for web files, including HTML, CSS, JavaScript, Markdown, LESS, Coffeescript, Handlebars, Angular, Ember, Knockdown, etc. Top links: Visual Studio 2013 content on the ASP.NET site are in the standard new releases area: http://www.asp.net/vnext ASP.NET and Web Tools for Visual Studio 2013 Release Notes Short intro videos on the new Visual Studio web editor features from Scott Hanselman and Mads Kristensen Announcing release of ASP.NET and Web Tools for Visual Studio 2013 post on the official .NET Web Development and Tools Blog Scott Guthrie's post: Announcing the Release of Visual Studio 2013 and Great Improvements to ASP.NET and Entity Framework Okay, for those of you who are still with me, let's dig in a bit. Quick web dev notes on downloading and installing Visual Studio 2013 I found Visual Studio 2013 to be a pretty fast install. According to Brian Harry's release post, installing over pre-release versions of Visual Studio is supported.  I've installed the release version over pre-release versions, and it worked fine. If you're only going to be doing web development, you can speed up the install if you just select Web Developer tools. Of course, as a good Microsoft employee, I'll mention that you might also want to install some of those other features, like the Store apps for Windows 8 and the Windows Phone 8.0 SDK, but they do download and install a lot of other stuff (e.g. the Windows Phone SDK sets up Hyper-V and downloads several GB's of VM's). So if you're planning just to do web development for now, you can pick just the Web Developer Tools and install the other stuff later. If you've got a fast internet connection, I recommend using the web installer instead of downloading the ISO. The ISO includes all the features, whereas the web installer just downloads what you're installing. Visual Studio 2013 development settings and color theme When you start up Visual Studio, it'll prompt you to pick some defaults. These are totally up to you -whatever suits your development style - and you can change them later. As I said, these are completely up to you. I recommend either the Web Development or Web Development (Code Only) settings. The only real difference is that Code Only hides the toolbars, and you can switch between them using Tools / Import and Export Settings / Reset. Web Development settings Web Development (code only) settings Usually I've just gone with Web Development (code only) in the past because I just want to focus on the code, although the Standard toolbar does make it easier to switch default web browsers. More on that later. Color theme Sigh. Okay, everyone's got their favorite colors. I alternate between Light and Dark depending on my mood, and I personally like how the low contrast on the window chrome in those themes puts the emphasis on my code rather than the tabs and toolbars. I know some people got pretty worked up over that, though, and wanted the blue theme back. I personally don't like it - it reminds me of ancient versions of Visual Studio that I don't want to think about anymore. So here's the thing: if you install Visual Studio Ultimate, it defaults to Blue. The other versions default to Light. If you use Blue, I won't criticize you - out loud, that is. You can change themes really easily - either Tools / Options / Environment / General, or the smart way: ctrl+q for quick launch, then type Theme and hit enter. Signing in During the first run, you'll be prompted to sign in. You don't have to - you can click the "Not now, maybe later" link at the bottom of that dialog. I recommend signing in, though. It's not hooked in with licensing or tracking the kind of code you write to sell you components. It is doing good things, like  syncing your Visual Studio settings between computers. More about that here. So, you don't have to, but I sure do. Overview of shiny new things in ASP.NET land There are a lot of good new things in ASP.NET. I'll list some of my favorite here, but you can read more on the ASP.NET site. One ASP.NET You've heard us talk about this for a while. The idea is that options are good, but choice can be a burden. When you start a new ASP.NET project, why should you have to make a tough decision - with long-term consequences - about how your application will work? If you want to use ASP.NET Web Forms, but have the option of adding in ASP.NET MVC later, why should that be hard? It's all ASP.NET, right? Ideally, you'd just decide that you want to use ASP.NET to build sites and services, and you could use the appropriate tools (the green blocks below) as you needed them. So, here it is. When you create a new ASP.NET application, you just create an ASP.NET application. Next, you can pick from some templates to get you started... but these are different. They're not "painful decision" templates, they're just some starting pieces. And, most importantly, you can mix and match. I can pick a "mostly" Web Forms template, but include MVC and Web API folders and core references. If you've tried to mix and match in the past, you're probably aware that it was possible, but not pleasant. ASP.NET MVC project files contained special project type GUIDs, so you'd only get controller scaffolding support in a Web Forms project if you manually edited the csproj file. Features in one stack didn't work in others. Project templates were painful choices. That's no longer the case. Hooray! I just did a demo in a presentation last week where I created a new Web Forms + MVC + Web API site, built a model, scaffolded MVC and Web API controllers with EF Code First, add data in the MVC view, viewed it in Web API, then added a GridView to the Web Forms Default.aspx page and bound it to the Model. In about 5 minutes. Sure, it's a simple example, but it's great to be able to share code and features across the whole ASP.NET family. Authentication In the past, authentication was built into the templates. So, for instance, there was an ASP.NET MVC 4 Intranet Project template which created a new ASP.NET MVC 4 application that was preconfigured for Windows Authentication. All of that authentication stuff was built into each template, so they varied between the stacks, and you couldn't reuse them. You didn't see a lot of changes to the authentication options, since they required big changes to a bunch of project templates. Now, the new project dialog includes a common authentication experience. When you hit the Change Authentication button, you get some common options that work the same way regardless of the template or reference settings you've made. These options work on all ASP.NET frameworks, and all hosting environments (IIS, IIS Express, or OWIN for self-host) The default is Individual User Accounts: This is the standard "create a local account, using username / password or OAuth" thing; however, it's all built on the new Identity system. More on that in a second. The one setting that has some configuration to it is Organizational Accounts, which lets you configure authentication using Active Directory, Windows Azure Active Directory, or Office 365. Identity There's a new identity system. We've taken the best parts of the previous ASP.NET Membership and Simple Identity systems, rolled in a lot of feedback and made big enhancements to support important developer concerns like unit testing and extensiblity. I've written long posts about ASP.NET identity, and I'll do it again. Soon. This is not that post. The short version is that I think we've finally got just the right Identity system. Some of my favorite features: There are simple, sensible defaults that work well - you can File / New / Run / Register / Login, and everything works. It supports standard username / password as well as external authentication (OAuth, etc.). It's easy to customize without having to re-implement an entire provider. It's built using pluggable pieces, rather than one large monolithic system. It's built using interfaces like IUser and IRole that allow for unit testing, dependency injection, etc. You can easily add user profile data (e.g. URL, twitter handle, birthday). You just add properties to your ApplicationUser model and they'll automatically be persisted. Complete control over how the identity data is persisted. By default, everything works with Entity Framework Code First, but it's built to support changes from small (modify the schema) to big (use another ORM, store your data in a document database or in the cloud or in XML or in the EXIF data of your desktop background or whatever). It's configured via OWIN. More on OWIN and Katana later, but the fact that it's built using OWIN means it's portable. You can find out more in the Authentication and Identity section of the ASP.NET site (and lots more content will be going up there soon). New Bootstrap based project templates The new project templates are built using Bootstrap 3. Bootstrap (formerly Twitter Bootstrap) is a front-end framework that brings a lot of nice benefits: It's responsive, so your projects will automatically scale to device width using CSS media queries. For example, menus are full size on a desktop browser, but on narrower screens you automatically get a mobile-friendly menu. The built-in Bootstrap styles make your standard page elements (headers, footers, buttons, form inputs, tables etc.) look nice and modern. Bootstrap is themeable, so you can reskin your whole site by dropping in a new Bootstrap theme. Since Bootstrap is pretty popular across the web development community, this gives you a large and rapidly growing variety of templates (free and paid) to choose from. Bootstrap also includes a lot of very useful things: components (like progress bars and badges), useful glyphicons, and some jQuery plugins for tooltips, dropdowns, carousels, etc.). Here's a look at how the responsive part works. When the page is full screen, the menu and header are optimized for a wide screen display: When I shrink the page down (this is all based on page width, not useragent sniffing) the menu turns into a nice mobile-friendly dropdown: For a quick example, I grabbed a new free theme off bootswatch.com. For simple themes, you just need to download the boostrap.css file and replace the /content/bootstrap.css file in your project. Now when I refresh the page, I've got a new theme: Scaffolding The big change in scaffolding is that it's one system that works across ASP.NET. You can create a new Empty Web project or Web Forms project and you'll get the Scaffold context menus. For release, we've got MVC 5 and Web API 2 controllers. We had a preview of Web Forms scaffolding in the preview releases, but they weren't fully baked for RTM. Look for them in a future update, expected pretty soon. This scaffolding system wasn't just changed to work across the ASP.NET frameworks, it's also built to enable future extensibility. That's not in this release, but should also hopefully be out soon. Project Readme page This is a small thing, but I really like it. When you create a new project, you get a Project_Readme.html page that's added to the root of your project and opens in the Visual Studio built-in browser. I love it. A long time ago, when you created a new project we just dumped it on you and left you scratching your head about what to do next. Not ideal. Then we started adding a bunch of Getting Started information to the new project templates. That told you what to do next, but you had to delete all of that stuff out of your website. It doesn't belong there. Not ideal. This is a simple HTML file that's not integrated into your project code at all. You can delete it if you want. But, it shows a lot of helpful links that are current for the project you just created. In the future, if we add new wacky project types, they can create readme docs with specific information on how to do appropriately wacky things. Side note: I really like that they used the internal browser in Visual Studio to show this content rather than popping open an HTML page in the default browser. I hate that. It's annoying. If you're doing that, I hope you'll stop. What if some unnamed person has 40 or 90 tabs saved in their browser session? When you pop open your "Thanks for installing my Visual Studio extension!" page, all eleventy billion tabs start up and I wish I'd never installed your thing. Be like these guys and pop stuff Visual Studio specific HTML docs in the Visual Studio browser. ASP.NET MVC 5 The biggest change with ASP.NET MVC 5 is that it's no longer a separate project type. It integrates well with the rest of ASP.NET. In addition to that and the other common features we've already looked at (Bootstrap templates, Identity, authentication), here's what's new for ASP.NET MVC. Attribute routing ASP.NET MVC now supports attribute routing, thanks to a contribution by Tim McCall, the author of http://attributerouting.net. With attribute routing you can specify your routes by annotating your actions and controllers. This supports some pretty complex, customized routing scenarios, and it allows you to keep your route information right with your controller actions if you'd like. Here's a controller that includes an action whose method name is Hiding, but I've used AttributeRouting to configure it to /spaghetti/with-nesting/where-is-waldo public class SampleController : Controller { [Route("spaghetti/with-nesting/where-is-waldo")] public string Hiding() { return "You found me!"; } } I enable that in my RouteConfig.cs, and I can use that in conjunction with my other MVC routes like this: public class RouteConfig { public static void RegisterRoutes(RouteCollection routes) { routes.IgnoreRoute("{resource}.axd/{*pathInfo}"); routes.MapMvcAttributeRoutes(); routes.MapRoute( name: "Default", url: "{controller}/{action}/{id}", defaults: new { controller = "Home", action = "Index", id = UrlParameter.Optional } ); } } You can read more about Attribute Routing in ASP.NET MVC 5 here. Filter enhancements There are two new additions to filters: Authentication Filters and Filter Overrides. Authentication filters are a new kind of filter in ASP.NET MVC that run prior to authorization filters in the ASP.NET MVC pipeline and allow you to specify authentication logic per-action, per-controller, or globally for all controllers. Authentication filters process credentials in the request and provide a corresponding principal. Authentication filters can also add authentication challenges in response to unauthorized requests. Override filters let you change which filters apply to a given action method or controller. Override filters specify a set of filter types that should not be run for a given scope (action or controller). This allows you to configure filters that apply globally but then exclude certain global filters from applying to specific actions or controllers. ASP.NET Web API 2 ASP.NET Web API 2 includes a lot of new features. Attribute Routing ASP.NET Web API supports the same attribute routing system that's in ASP.NET MVC 5. You can read more about the Attribute Routing features in Web API in this article. OAuth 2.0 ASP.NET Web API picks up OAuth 2.0 support, using security middleware running on OWIN (discussed below). This is great for features like authenticated Single Page Applications. OData Improvements ASP.NET Web API now has full OData support. That required adding in some of the most powerful operators: $select, $expand, $batch and $value. You can read more about OData operator support in this article by Mike Wasson. Lots more There's a huge list of other features, including CORS (cross-origin request sharing), IHttpActionResult, IHttpRequestContext, and more. I think the best overview is in the release notes. OWIN and Katana I've written about OWIN and Katana recently. I'm a big fan. OWIN is the Open Web Interfaces for .NET. It's a spec, like HTML or HTTP, so you can't install OWIN. The benefit of OWIN is that it's a community specification, so anyone who implements it can plug into the ASP.NET stack, either as middleware or as a host. Katana is the Microsoft implementation of OWIN. It leverages OWIN to wire up things like authentication, handlers, modules, IIS hosting, etc., so ASP.NET can host OWIN components and Katana components can run in someone else's OWIN implementation. Howard Dierking just wrote a cool article in MSDN magazine describing Katana in depth: Getting Started with the Katana Project. He had an interesting example showing an OWIN based pipeline which leveraged SignalR, ASP.NET Web API and NancyFx components in the same stack. If this kind of thing makes sense to you, that's great. If it doesn't, don't worry, but keep an eye on it. You're going to see some cool things happen as a result of ASP.NET becoming more and more pluggable. Visual Studio Web Tools Okay, this stuff's just crazy. Visual Studio has been adding some nice web dev features over the past few years, but they've really cranked it up for this release. Visual Studio is by far my favorite code editor for all web files: CSS, HTML, JavaScript, and lots of popular libraries. Stop thinking of Visual Studio as a big editor that you only use to write back-end code. Stop editing HTML and CSS in Notepad (or Sublime, Notepad++, etc.). Visual Studio starts up in under 2 seconds on a modern computer with an SSD. Misspelling HTML attributes or your CSS classes or jQuery or Angular syntax is stupid. It doesn't make you a better developer, it makes you a silly person who wastes time. Browser Link Browser Link is a real-time, two-way connection between Visual Studio and all connected browsers. It's only attached when you're running locally, in debug, but it applies to any and all connected browser, including emulators. You may have seen demos that showed the browsers refreshing based on changes in the editor, and I'll agree that's pretty cool. But it's really just the start. It's a two-way connection, and it's built for extensiblity. That means you can write extensions that push information from your running application (in IE, Chrome, a mobile emulator, etc.) back to Visual Studio. Mads and team have showed off some demonstrations where they enabled edit mode in the browser which updated the source HTML back on the browser. It's also possible to look at how the rendered HTML performs, check for compatibility issues, watch for unused CSS classes, the sky's the limit. New HTML editor The previous HTML editor had a lot of old code that didn't allow for improvements. The team rewrote the HTML editor to take advantage of the new(ish) extensibility features in Visual Studio, which then allowed them to add in all kinds of features - things like CSS Class and ID IntelliSense (so you type style="" and get a list of classes and ID's for your project), smart indent based on how your document is formatted, JavaScript reference auto-sync, etc. Here's a 3 minute tour from Mads Kristensen. The previous HTML editor had a lot of old code that didn't allow for improvements. The team rewrote the HTML editor to take advantage of the new(ish) extensibility features in Visual Studio, which then allowed them to add in all kinds of features - things like CSS Class and ID IntelliSense (so you type style="" and get a list of classes and ID's for your project), smart indent based on how your document is formatted, JavaScript reference auto-sync, etc. Lots more Visual Studio web dev features That's just a sampling - there's a ton of great features for JavaScript editing, CSS editing, publishing, and Page Inspector (which shows real-time rendering of your page inside Visual Studio). Here are some more short videos showing those features. Lots, lots more Okay, that's just a summary, and it's still quite a bit. Head on over to http://asp.net/vnext for more information, and download Visual Studio 2013 now to get started!

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  • SQL SERVER – Securing TRUNCATE Permissions in SQL Server

    - by pinaldave
    Download the Script of this article from here. On December 11, 2010, Vinod Kumar, a Databases & BI technology evangelist from Microsoft Corporation, graced Ahmedabad by spending some time with the Community during the Community Tech Days (CTD) event. As he was running through a few demos, Vinod asked the audience one of the most fundamental and common interview questions – “What is the difference between a DELETE and TRUNCATE?“ Ahmedabad SQL Server User Group Expert Nakul Vachhrajani has come up with excellent solutions of the same. I must congratulate Nakul for this excellent solution and as a encouragement to User Group member, I am publishing the same article over here. Nakul Vachhrajani is a Software Specialist and systems development professional with Patni Computer Systems Limited. He has functional experience spanning legacy code deprecation, system design, documentation, development, implementation, testing, maintenance and support of complex systems, providing business intelligence solutions, database administration, performance tuning, optimization, product management, release engineering, process definition and implementation. He has comprehensive grasp on Database Administration, Development and Implementation with MS SQL Server and C, C++, Visual C++/C#. He has about 6 years of total experience in information technology. Nakul is an member of the Ahmedabad and Gandhinagar SQL Server User Groups, and actively contributes to the community by actively participating in multiple forums and websites like SQLAuthority.com, BeyondRelational.com, SQLServerCentral.com and many others. Please note: The opinions expressed herein are Nakul own personal opinions and do not represent his employer’s view in anyway. All data from everywhere here on Earth go through a series of  four distinct operations, identified by the words: CREATE, READ, UPDATE and DELETE, or simply, CRUD. Putting in Microsoft SQL Server terms, is the process goes like this: INSERT, SELECT, UPDATE and DELETE/TRUNCATE. Quite a few interesting responses were received and evaluated live during the session. To summarize them, the most important similarity that came out was that both DELETE and TRUNCATE participate in transactions. The major differences (not all) that came out of the exercise were: DELETE: DELETE supports a WHERE clause DELETE removes rows from a table, row-by-row Because DELETE moves row-by-row, it acquires a row-level lock Depending upon the recovery model of the database, DELETE is a fully-logged operation. Because DELETE moves row-by-row, it can fire off triggers TRUNCATE: TRUNCATE does not support a WHERE clause TRUNCATE works by directly removing the individual data pages of a table TRUNCATE directly occupies a table-level lock. (Because a lock is acquired, and because TRUNCATE can also participate in a transaction, it has to be a logged operation) TRUNCATE is, therefore, a minimally-logged operation; again, this depends upon the recovery model of the database Triggers are not fired when TRUNCATE is used (because individual row deletions are not logged) Finally, Vinod popped the big homework question that must be critically analyzed: “We know that we can restrict a DELETE operation to a particular user, but how can we restrict the TRUNCATE operation to a particular user?” After returning home and having a nice cup of coffee, I noticed that my gray cells immediately started to work. Below was the result of my research. As what is always said, the devil is in the details. Upon looking at the Permissions section for the TRUNCATE statement in Books On Line, the following jumps right out: “The minimum permission required is ALTER on table_name. TRUNCATE TABLE permissions default to the table owner, members of the sysadmin fixed server role, and the db_owner and db_ddladmin fixed database roles, and are not transferable. However, you can incorporate the TRUNCATE TABLE statement within a module, such as a stored procedure, and grant appropriate permissions to the module using the EXECUTE AS clause.“ Now, what does this mean? Unlike DELETE, one cannot directly assign permissions to a user/set of users allowing or revoking TRUNCATE rights. However, there is a way to circumvent this. It is important to recall that in Microsoft SQL Server, database engine security surrounds the concept of a “securable”, which is any object like a table, stored procedure, trigger, etc. Rights are assigned to a principal on a securable. Refer to the image below (taken from the SQL Server Books On Line). urable”, which is any object like a table, stored procedure, trigger, etc. Rights are assigned to a principal on a securable. Refer to the image below (taken from the SQL Server Books On Line). SETTING UP THE ENVIRONMENT – (01A_Truncate Table Permissions.sql) Script Provided at the end of the article. By the end of this demo, one will be able to do all the CRUD operations, except the TRUNCATE, and the other will only be able to execute the TRUNCATE. All you will need for this test is any edition of SQL Server 2008. (With minor changes, these scripts can be made to work with SQL 2005.) We begin by creating the following: 1.       A test database 2.        Two database roles: associated logins and users 3.       Switch over to the test database and create a test table. Then, add some data into it. I am using row constructors, which is new to SQL 2008. Creating the modules that will be used to enforce permissions 1.       We have already created one of the modules that we will be assigning permissions to. That module is the table: TruncatePermissionsTest 2.       We will now create two stored procedures; one is for the DELETE operation and the other for the TRUNCATE operation. Please note that for all practical purposes, the end result is the same – all data from the table TruncatePermissionsTest is removed Assigning the permissions Now comes the most important part of the demonstration – assigning permissions. A permissions matrix can be worked out as under: To apply the security rights, we use the GRANT and DENY clauses, as under: That’s it! We are now ready for our big test! THE TEST (01B_Truncate Table Test Queries.sql) Script Provided at the end of the article. I will now need two separate SSMS connections, one with the login AllowedTruncate and the other with the login RestrictedTruncate. Running the test is simple; all that’s required is to run through the script – 01B_Truncate Table Test Queries.sql. What I will demonstrate here via screen-shots is the behavior of SQL Server when logged in as the AllowedTruncate user. There are a few other combinations than what are highlighted here. I will leave the reader the right to explore the behavior of the RestrictedTruncate user and these additional scenarios, as a form of self-study. 1.       Testing SELECT permissions 2.       Testing TRUNCATE permissions (Remember, “deny by default”?) 3.       Trying to circumvent security by trying to TRUNCATE the table using the stored procedure Hence, we have now proved that a user can indeed be assigned permissions to specifically assign TRUNCATE permissions. I also hope that the above has sparked curiosity towards putting some security around the probably “destructive” operations of DELETE and TRUNCATE. I would like to wish each and every one of the readers a very happy and secure time with Microsoft SQL Server. (Please find the scripts – 01A_Truncate Table Permissions.sql and 01B_Truncate Table Test Queries.sql that have been used in this demonstration. Please note that these scripts contain purely test-level code only. These scripts must not, at any cost, be used in the reader’s production environments). 01A_Truncate Table Permissions.sql /* ***************************************************************************************************************** Developed By          : Nakul Vachhrajani Functionality         : This demo is focused on how to allow only TRUNCATE permissions to a particular user How to Use            : 1. Run through, step-by-step through the sequence till Step 08 to create a test database 2. Switch over to the "Truncate Table Test Queries.sql" and execute it step-by-step in two different SSMS windows, one where you have logged in as 'RestrictedTruncate', and the other as 'AllowedTruncate' 3. Come back to "Truncate Table Permissions.sql" 4. Execute Step 10 to cleanup! Modifications         : December 13, 2010 - NAV - Updated to add a security matrix and improve code readability when applying security December 12, 2010 - NAV - Created ***************************************************************************************************************** */ -- Step 01: Create a new test database CREATE DATABASE TruncateTestDB GO USE TruncateTestDB GO -- Step 02: Add roles and users to demonstrate the security of the Truncate operation -- 2a. Create the new roles CREATE ROLE AllowedTruncateRole; GO CREATE ROLE RestrictedTruncateRole; GO -- 2b. Create new logins CREATE LOGIN AllowedTruncate WITH PASSWORD = 'truncate@2010', CHECK_POLICY = ON GO CREATE LOGIN RestrictedTruncate WITH PASSWORD = 'truncate@2010', CHECK_POLICY = ON GO -- 2c. Create new Users using the roles and logins created aboave CREATE USER TruncateUser FOR LOGIN AllowedTruncate WITH DEFAULT_SCHEMA = dbo GO CREATE USER NoTruncateUser FOR LOGIN RestrictedTruncate WITH DEFAULT_SCHEMA = dbo GO -- 2d. Add the newly created login to the newly created role sp_addrolemember 'AllowedTruncateRole','TruncateUser' GO sp_addrolemember 'RestrictedTruncateRole','NoTruncateUser' GO -- Step 03: Change over to the test database USE TruncateTestDB GO -- Step 04: Create a test table within the test databse CREATE TABLE TruncatePermissionsTest (Id INT IDENTITY(1,1), Name NVARCHAR(50)) GO -- Step 05: Populate the required data INSERT INTO TruncatePermissionsTest VALUES (N'Delhi'), (N'Mumbai'), (N'Ahmedabad') GO -- Step 06: Encapsulate the DELETE within another module CREATE PROCEDURE proc_DeleteMyTable WITH EXECUTE AS SELF AS DELETE FROM TruncateTestDB..TruncatePermissionsTest GO -- Step 07: Encapsulate the TRUNCATE within another module CREATE PROCEDURE proc_TruncateMyTable WITH EXECUTE AS SELF AS TRUNCATE TABLE TruncateTestDB..TruncatePermissionsTest GO -- Step 08: Apply Security /* *****************************SECURITY MATRIX*************************************** =================================================================================== Object                   | Permissions |                 Login |             | AllowedTruncate   |   RestrictedTruncate |             |User:NoTruncateUser|   User:TruncateUser =================================================================================== TruncatePermissionsTest  | SELECT,     |      GRANT        |      (Default) | INSERT,     |                   | | UPDATE,     |                   | | DELETE      |                   | -------------------------+-------------+-------------------+----------------------- TruncatePermissionsTest  | ALTER       |      DENY         |      (Default) -------------------------+-------------+----*/----------------+----------------------- proc_DeleteMyTable | EXECUTE | GRANT | DENY -------------------------+-------------+-------------------+----------------------- proc_TruncateMyTable | EXECUTE | DENY | GRANT -------------------------+-------------+-------------------+----------------------- *****************************SECURITY MATRIX*************************************** */ /* Table: TruncatePermissionsTest*/ GRANT SELECT, INSERT, UPDATE, DELETE ON TruncateTestDB..TruncatePermissionsTest TO NoTruncateUser GO DENY ALTER ON TruncateTestDB..TruncatePermissionsTest TO NoTruncateUser GO /* Procedure: proc_DeleteMyTable*/ GRANT EXECUTE ON TruncateTestDB..proc_DeleteMyTable TO NoTruncateUser GO DENY EXECUTE ON TruncateTestDB..proc_DeleteMyTable TO TruncateUser GO /* Procedure: proc_TruncateMyTable*/ DENY EXECUTE ON TruncateTestDB..proc_TruncateMyTable TO NoTruncateUser GO GRANT EXECUTE ON TruncateTestDB..proc_TruncateMyTable TO TruncateUser GO -- Step 09: Test --Switch over to the "Truncate Table Test Queries.sql" and execute it step-by-step in two different SSMS windows: --    1. one where you have logged in as 'RestrictedTruncate', and --    2. the other as 'AllowedTruncate' -- Step 10: Cleanup sp_droprolemember 'AllowedTruncateRole','TruncateUser' GO sp_droprolemember 'RestrictedTruncateRole','NoTruncateUser' GO DROP USER TruncateUser GO DROP USER NoTruncateUser GO DROP LOGIN AllowedTruncate GO DROP LOGIN RestrictedTruncate GO DROP ROLE AllowedTruncateRole GO DROP ROLE RestrictedTruncateRole GO USE MASTER GO DROP DATABASE TruncateTestDB GO 01B_Truncate Table Test Queries.sql /* ***************************************************************************************************************** Developed By          : Nakul Vachhrajani Functionality         : This demo is focused on how to allow only TRUNCATE permissions to a particular user How to Use            : 1. Switch over to this from "Truncate Table Permissions.sql", Step #09 2. Execute this step-by-step in two different SSMS windows a. One where you have logged in as 'RestrictedTruncate', and b. The other as 'AllowedTruncate' 3. Return back to "Truncate Table Permissions.sql" 4. Execute Step 10 to cleanup! Modifications         : December 12, 2010 - NAV - Created ***************************************************************************************************************** */ -- Step 09A: Switch to the test database USE TruncateTestDB GO -- Step 09B: Ensure that we have valid data SELECT * FROM TruncatePermissionsTest GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 1 -- The SELECT permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. --Step 09C: Attempt to Truncate Data from the table without using the stored procedure TRUNCATE TABLE TruncatePermissionsTest GO -- (Expected: Following error will occur) --  Msg 1088, Level 16, State 7, Line 2 --  Cannot find the object "TruncatePermissionsTest" because it does not exist or you do not have permissions. -- Step 09D:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'London'), (N'Paris'), (N'Berlin') GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 1 -- The INSERT permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. --Step 09E: Attempt to Truncate Data from the table using the stored procedure EXEC proc_TruncateMyTable GO -- (Expected: Will execute successfully with 'AllowedTruncate' user, will error out as under with 'RestrictedTruncate') -- Msg 229, Level 14, State 5, Procedure proc_TruncateMyTable, Line 1 -- The EXECUTE permission was denied on the object 'proc_TruncateMyTable', database 'TruncateTestDB', schema 'dbo'. -- Step 09F:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'Madrid'), (N'Rome'), (N'Athens') GO --Step 09G: Attempt to Delete Data from the table without using the stored procedure DELETE FROM TruncatePermissionsTest GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Line 2 -- The DELETE permission was denied on the object 'TruncatePermissionsTest', database 'TruncateTestDB', schema 'dbo'. -- Step 09H:Regenerate Test Data INSERT INTO TruncatePermissionsTest VALUES (N'Spain'), (N'Italy'), (N'Greece') GO --Step 09I: Attempt to Delete Data from the table using the stored procedure EXEC proc_DeleteMyTable GO -- (Expected: Following error will occur if logged in as "AllowedTruncate") -- Msg 229, Level 14, State 5, Procedure proc_DeleteMyTable, Line 1 -- The EXECUTE permission was denied on the object 'proc_DeleteMyTable', database 'TruncateTestDB', schema 'dbo'. --Step 09J: Close this SSMS window and return back to "Truncate Table Permissions.sql" Thank you Nakul to take up the challenge and prove that Ahmedabad and Gandhinagar SQL Server User Group has talent to solve difficult problems. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Best Practices, Pinal Dave, Readers Contribution, Readers Question, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Security, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Rendering ASP.NET MVC Views to String

    - by Rick Strahl
    It's not uncommon in my applications that I require longish text output that does not have to be rendered into the HTTP output stream. The most common scenario I have for 'template driven' non-Web text is for emails of all sorts. Logon confirmations and verifications, email confirmations for things like orders, status updates or scheduler notifications - all of which require merged text output both within and sometimes outside of Web applications. On other occasions I also need to capture the output from certain views for logging purposes. Rather than creating text output in code, it's much nicer to use the rendering mechanism that ASP.NET MVC already provides by way of it's ViewEngines - using Razor or WebForms views - to render output to a string. This is nice because it uses the same familiar rendering mechanism that I already use for my HTTP output and it also solves the problem of where to store the templates for rendering this content in nothing more than perhaps a separate view folder. The good news is that ASP.NET MVC's rendering engine is much more modular than the full ASP.NET runtime engine which was a real pain in the butt to coerce into rendering output to string. With MVC the rendering engine has been separated out from core ASP.NET runtime, so it's actually a lot easier to get View output into a string. Getting View Output from within an MVC Application If you need to generate string output from an MVC and pass some model data to it, the process to capture this output is fairly straight forward and involves only a handful of lines of code. The catch is that this particular approach requires that you have an active ControllerContext that can be passed to the view. This means that the following approach is limited to access from within Controller methods. Here's a class that wraps the process and provides both instance and static methods to handle the rendering:/// <summary> /// Class that renders MVC views to a string using the /// standard MVC View Engine to render the view. /// /// Note: This class can only be used within MVC /// applications that have an active ControllerContext. /// </summary> public class ViewRenderer { /// <summary> /// Required Controller Context /// </summary> protected ControllerContext Context { get; set; } public ViewRenderer(ControllerContext controllerContext) { Context = controllerContext; } /// <summary> /// Renders a full MVC view to a string. Will render with the full MVC /// View engine including running _ViewStart and merging into _Layout /// </summary> /// <param name="viewPath"> /// The path to the view to render. Either in same controller, shared by /// name or as fully qualified ~/ path including extension /// </param> /// <param name="model">The model to render the view with</param> /// <returns>String of the rendered view or null on error</returns> public string RenderView(string viewPath, object model) { return RenderViewToStringInternal(viewPath, model, false); } /// <summary> /// Renders a partial MVC view to string. Use this method to render /// a partial view that doesn't merge with _Layout and doesn't fire /// _ViewStart. /// </summary> /// <param name="viewPath"> /// The path to the view to render. Either in same controller, shared by /// name or as fully qualified ~/ path including extension /// </param> /// <param name="model">The model to pass to the viewRenderer</param> /// <returns>String of the rendered view or null on error</returns> public string RenderPartialView(string viewPath, object model) { return RenderViewToStringInternal(viewPath, model, true); } public static string RenderView(string viewPath, object model, ControllerContext controllerContext) { ViewRenderer renderer = new ViewRenderer(controllerContext); return renderer.RenderView(viewPath, model); } public static string RenderPartialView(string viewPath, object model, ControllerContext controllerContext) { ViewRenderer renderer = new ViewRenderer(controllerContext); return renderer.RenderPartialView(viewPath, model); } protected string RenderViewToStringInternal(string viewPath, object model, bool partial = false) { // first find the ViewEngine for this view ViewEngineResult viewEngineResult = null; if (partial) viewEngineResult = ViewEngines.Engines.FindPartialView(Context, viewPath); else viewEngineResult = ViewEngines.Engines.FindView(Context, viewPath, null); if (viewEngineResult == null) throw new FileNotFoundException(Properties.Resources.ViewCouldNotBeFound); // get the view and attach the model to view data var view = viewEngineResult.View; Context.Controller.ViewData.Model = model; string result = null; using (var sw = new StringWriter()) { var ctx = new ViewContext(Context, view, Context.Controller.ViewData, Context.Controller.TempData, sw); view.Render(ctx, sw); result = sw.ToString(); } return result; } } The key is the RenderViewToStringInternal method. The method first tries to find the view to render based on its path which can either be in the current controller's view path or the shared view path using its simple name (PasswordRecovery) or alternately by its full virtual path (~/Views/Templates/PasswordRecovery.cshtml). This code should work both for Razor and WebForms views although I've only tried it with Razor Views. Note that WebForms Views might actually be better for plain text as Razor adds all sorts of white space into its output when there are code blocks in the template. The Web Forms engine provides more accurate rendering for raw text scenarios. Once a view engine is found the view to render can be retrieved. Views in MVC render based on data that comes off the controller like the ViewData which contains the model along with the actual ViewData and ViewBag. From the View and some of the Context data a ViewContext is created which is then used to render the view with. The View picks up the Model and other data from the ViewContext internally and processes the View the same it would be processed if it were to send its output into the HTTP output stream. The difference is that we can override the ViewContext's output stream which we provide and capture into a StringWriter(). After rendering completes the result holds the output string. If an error occurs the error behavior is similar what you see with regular MVC errors - you get a full yellow screen of death including the view error information with the line of error highlighted. It's your responsibility to handle the error - or let it bubble up to your regular Controller Error filter if you have one. To use the simple class you only need a single line of code if you call the static methods. Here's an example of some Controller code that is used to send a user notification to a customer via email in one of my applications:[HttpPost] public ActionResult ContactSeller(ContactSellerViewModel model) { InitializeViewModel(model); var entryBus = new busEntry(); var entry = entryBus.LoadByDisplayId(model.EntryId); if ( string.IsNullOrEmpty(model.Email) ) entryBus.ValidationErrors.Add("Email address can't be empty.","Email"); if ( string.IsNullOrEmpty(model.Message)) entryBus.ValidationErrors.Add("Message can't be empty.","Message"); model.EntryId = entry.DisplayId; model.EntryTitle = entry.Title; if (entryBus.ValidationErrors.Count > 0) { ErrorDisplay.AddMessages(entryBus.ValidationErrors); ErrorDisplay.ShowError("Please correct the following:"); } else { string message = ViewRenderer.RenderView("~/views/template/ContactSellerEmail.cshtml",model, ControllerContext); string title = entry.Title + " (" + entry.DisplayId + ") - " + App.Configuration.ApplicationName; AppUtils.SendEmail(title, message, model.Email, entry.User.Email, false, false)) } return View(model); } Simple! The view in this case is just a plain MVC view and in this case it's a very simple plain text email message (edited for brevity here) that is created and sent off:@model ContactSellerViewModel @{ Layout = null; }re: @Model.EntryTitle @Model.ListingUrl @Model.Message ** SECURITY ADVISORY - AVOID SCAMS ** Avoid: wiring money, cross-border deals, work-at-home ** Beware: cashier checks, money orders, escrow, shipping ** More Info: @(App.Configuration.ApplicationBaseUrl)scams.html Obviously this is a very simple view (I edited out more from this page to keep it brief) -  but other template views are much more complex HTML documents or long messages that are occasionally updated and they are a perfect fit for Razor rendering. It even works with nested partial views and _layout pages. Partial Rendering Notice that I'm rendering a full View here. In the view I explicitly set the Layout=null to avoid pulling in _layout.cshtml for this view. This can also be controlled externally by calling the RenderPartial method instead: string message = ViewRenderer.RenderPartialView("~/views/template/ContactSellerEmail.cshtml",model, ControllerContext); with this line of code no layout page (or _viewstart) will be loaded, so the output generated is just what's in the view. I find myself using Partials most of the time when rendering templates, since the target of templates usually tend to be emails or other HTML fragment like output, so the RenderPartialView() method is definitely useful to me. Rendering without a ControllerContext The preceding class is great when you're need template rendering from within MVC controller actions or anywhere where you have access to the request Controller. But if you don't have a controller context handy - maybe inside a utility function that is static, a non-Web application, or an operation that runs asynchronously in ASP.NET - which makes using the above code impossible. I haven't found a way to manually create a Controller context to provide the ViewContext() what it needs from outside of the MVC infrastructure. However, there are ways to accomplish this,  but they are a bit more complex. It's possible to host the RazorEngine on your own, which side steps all of the MVC framework and HTTP and just deals with the raw rendering engine. I wrote about this process in Hosting the Razor Engine in Non-Web Applications a long while back. It's quite a process to create a custom Razor engine and runtime, but it allows for all sorts of flexibility. There's also a RazorEngine CodePlex project that does something similar. I've been meaning to check out the latter but haven't gotten around to it since I have my own code to do this. The trick to hosting the RazorEngine to have it behave properly inside of an ASP.NET application and properly cache content so templates aren't constantly rebuild and reparsed. Anyway, in the same app as above I have one scenario where no ControllerContext is available: I have a background scheduler running inside of the app that fires on timed intervals. This process could be external but because it's lightweight we decided to fire it right inside of the ASP.NET app on a separate thread. In my app the code that renders these templates does something like this:var model = new SearchNotificationViewModel() { Entries = entries, Notification = notification, User = user }; // TODO: Need logging for errors sending string razorError = null; var result = AppUtils.RenderRazorTemplate("~/views/template/SearchNotificationTemplate.cshtml", model, razorError); which references a couple of helper functions that set up my RazorFolderHostContainer class:public static string RenderRazorTemplate(string virtualPath, object model,string errorMessage = null) { var razor = AppUtils.CreateRazorHost(); var path = virtualPath.Replace("~/", "").Replace("~", "").Replace("/", "\\"); var merged = razor.RenderTemplateToString(path, model); if (merged == null) errorMessage = razor.ErrorMessage; return merged; } /// <summary> /// Creates a RazorStringHostContainer and starts it /// Call .Stop() when you're done with it. /// /// This is a static instance /// </summary> /// <param name="virtualPath"></param> /// <param name="binBasePath"></param> /// <param name="forceLoad"></param> /// <returns></returns> public static RazorFolderHostContainer CreateRazorHost(string binBasePath = null, bool forceLoad = false) { if (binBasePath == null) { if (HttpContext.Current != null) binBasePath = HttpContext.Current.Server.MapPath("~/"); else binBasePath = AppDomain.CurrentDomain.BaseDirectory; } if (_RazorHost == null || forceLoad) { if (!binBasePath.EndsWith("\\")) binBasePath += "\\"; //var razor = new RazorStringHostContainer(); var razor = new RazorFolderHostContainer(); razor.TemplatePath = binBasePath; binBasePath += "bin\\"; razor.BaseBinaryFolder = binBasePath; razor.UseAppDomain = false; razor.ReferencedAssemblies.Add(binBasePath + "ClassifiedsBusiness.dll"); razor.ReferencedAssemblies.Add(binBasePath + "ClassifiedsWeb.dll"); razor.ReferencedAssemblies.Add(binBasePath + "Westwind.Utilities.dll"); razor.ReferencedAssemblies.Add(binBasePath + "Westwind.Web.dll"); razor.ReferencedAssemblies.Add(binBasePath + "Westwind.Web.Mvc.dll"); razor.ReferencedAssemblies.Add("System.Web.dll"); razor.ReferencedNamespaces.Add("System.Web"); razor.ReferencedNamespaces.Add("ClassifiedsBusiness"); razor.ReferencedNamespaces.Add("ClassifiedsWeb"); razor.ReferencedNamespaces.Add("Westwind.Web"); razor.ReferencedNamespaces.Add("Westwind.Utilities"); _RazorHost = razor; _RazorHost.Start(); //_RazorHost.Engine.Configuration.CompileToMemory = false; } return _RazorHost; } The RazorFolderHostContainer essentially is a full runtime that mimics a folder structure like a typical Web app does including caching semantics and compiling code only if code changes on disk. It maps a folder hierarchy to views using the ~/ path syntax. The host is then configured to add assemblies and namespaces. Unfortunately the engine is not exactly like MVC's Razor - the expression expansion and code execution are the same, but some of the support methods like sections, helpers etc. are not all there so templates have to be a bit simpler. There are other folder hosts provided as well to directly execute templates from strings (using RazorStringHostContainer). The following is an example of an HTML email template @inherits RazorHosting.RazorTemplateFolderHost <ClassifiedsWeb.SearchNotificationViewModel> <html> <head> <title>Search Notifications</title> <style> body { margin: 5px;font-family: Verdana, Arial; font-size: 10pt;} h3 { color: SteelBlue; } .entry-item { border-bottom: 1px solid grey; padding: 8px; margin-bottom: 5px; } </style> </head> <body> Hello @Model.User.Name,<br /> <p>Below are your Search Results for the search phrase:</p> <h3>@Model.Notification.SearchPhrase</h3> <small>since @TimeUtils.ShortDateString(Model.Notification.LastSearch)</small> <hr /> You can see that the syntax is a little different. Instead of the familiar @model header the raw Razor  @inherits tag is used to specify the template base class (which you can extend). I took a quick look through the feature set of RazorEngine on CodePlex (now Github I guess) and the template implementation they use is closer to MVC's razor but there are other differences. In the end don't expect exact behavior like MVC templates if you use an external Razor rendering engine. This is not what I would consider an ideal solution, but it works well enough for this project. My biggest concern is the overhead of hosting a second razor engine in a Web app and the fact that here the differences in template rendering between 'real' MVC Razor views and another RazorEngine really are noticeable. You win some, you lose some It's extremely nice to see that if you have a ControllerContext handy (which probably addresses 99% of Web app scenarios) rendering a view to string using the native MVC Razor engine is pretty simple. Kudos on making that happen - as it solves a problem I see in just about every Web application I work on. But it is a bummer that a ControllerContext is required to make this simple code work. It'd be really sweet if there was a way to render views without being so closely coupled to the ASP.NET or MVC infrastructure that requires a ControllerContext. Alternately it'd be nice to have a way for an MVC based application to create a minimal ControllerContext from scratch - maybe somebody's been down that path. I tried for a few hours to come up with a way to make that work but gave up in the soup of nested contexts (MVC/Controller/View/Http). I suspect going down this path would be similar to hosting the ASP.NET runtime requiring a WorkerRequest. Brrr…. The sad part is that it seems to me that a View should really not require much 'context' of any kind to render output to string. Yes there are a few things that clearly are required like paths to the virtual and possibly the disk paths to the root of the app, but beyond that view rendering should not require much. But, no such luck. For now custom RazorHosting seems to be the only way to make Razor rendering go outside of the MVC context… Resources Full ViewRenderer.cs source code from Westwind.Web.Mvc library Hosting the Razor Engine for Non-Web Applications RazorEngine on GitHub© Rick Strahl, West Wind Technologies, 2005-2012Posted in ASP.NET   ASP.NET  MVC   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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