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  • 8bpp Bitmap format on the Compact Framework

    - by Kieran
    Hi friends. I am messing around with Conway's Game of Life - http://en.wikipedia.org/wiki/Conway's_Game_of_Life I started out coding algorithmns for winforms and now want to port my work onto windows mobile 6.1 (compact framework). I came across an article by Jon Skeet where he compared several different algorithmns for calculating next generations in the game. He used an array of bytes to store a cells state (alive or dead) and then he would copy this array to an 8bpp bitmap. For each new generation, he works out the state of each byte, then copies the array to a bitmap, then draws that bitmap to a picturebox. void CreateInitialImage() { bitmap = new Bitmap(Width, Height, PixelFormat.Format8bppIndexed); ColorPalette palette = bitmap.Palette; palette.Entries[0] = Color.Black; palette.Entries[1] = Color.White; bitmap.Palette = palette; } public Image Render() { Rectangle rect = new Rectangle(0, 0, Width, Height); BitmapData bmpData = bitmap.LockBits(rect, ImageLockMode.ReadWrite, bitmap.PixelFormat); Marshal.Copy(Data, 0, bmpData.Scan0, Data.Length); bitmap.UnlockBits(bmpData); return bitmap; } His code above is beautifully simple and very fast to render. Jon is using Windows Forms but now I want to port my own version of this onto Windows Mobile 6.1 (Compact Framework) but . . . .there is no way to format a bitmap to 8bpp in the cf. Can anyone suggest a way of rendering an array of bytes to a drawable image in the CF. This array is created in code on the fly (it is NOT loaded from an image file on disk). I basically need to store an array of cells represented by bytes, they are either alive or dead and I then need to draw that array as an image. The game is particularly slow on the CF so I need to implement clever optimised algoritmns but also need to render as fast as possible and the above solution would be pretty dam perfect if only it was available on the compact framework. Many thanks for any help Any suggestions?

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  • Cannot figure out how to take in generic parameters for an Enterprise Framework library sql statemen

    - by KallDrexx
    I have written a specialized class to wrap up the enterprise library database functionality for easier usage. The reasoning for using the Enterprise Library is because my applications commonly connect to both oracle and sql server database systems. My wrapper handles both creating connection strings on the fly, connecting, and executing queries allowing my main code to only have to write a few lines of code to do database stuff and deal with error handling. As an example my ExecuteNonQuery method has the following declaration: /// <summary> /// Executes a query that returns no results (e.g. insert or update statements) /// </summary> /// <param name="sqlQuery"></param> /// <param name="parameters">Hashtable containing all the parameters for the query</param> /// <returns>The total number of records modified, -1 if an error occurred </returns> public int ExecuteNonQuery(string sqlQuery, Hashtable parameters) { // Make sure we are connected to the database if (!IsConnected) { ErrorHandler("Attempted to run a query without being connected to a database.", ErrorSeverity.Critical); return -1; } // Form the command DbCommand dbCommand = _database.GetSqlStringCommand(sqlQuery); // Add all the paramters foreach (string key in parameters.Keys) { if (parameters[key] == null) _database.AddInParameter(dbCommand, key, DbType.Object, null); else _database.AddInParameter(dbCommand, key, DbType.Object, parameters[key].ToString()); } return _database.ExecuteNonQuery(dbCommand); } _database is defined as private Database _database;. Hashtable parameters are created via code similar to p.Add("@param", value);. the issue I am having is that it seems that with enterprise library database framework you must declare the dbType of each parameter. This isn't an issue when you are calling the database code directly when forming the paramters but doesn't work for creating a generic abstraction class such as I have. In order to try and get around that I thought I could just use DbType.Object and figure the DB will figure it out based on the columns the sql is working with. Unfortunately, this is not the case as I get the following error: Implicit conversion from data type sql_variant to varchar is not allowed. Use the CONVERT function to run this query Is there any way to use generic parameters in a wrapper class or am I just going to have to move all my DB code into my main classes?

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  • Trying to understand the usage of class_eval

    - by eMxyzptlk
    Hello everyone, I'm using the rails-settings gem, and I'm trying to understand how you add functions to ActiveRecord classes (I'm building my own library for card games), and I noticed that this gem uses one of the Meta-programming techniques to add the function to the ActiveRecord::Base class (I'm far from Meta-programming master in ruby, but I'm trying to learn it) module RailsSettings class Railtie < Rails::Railtie initializer 'rails_settings.initialize', :after => :after_initialize do Railtie.extend_active_record end end class Railtie def self.extend_active_record ActiveRecord::Base.class_eval do def self.has_settings class_eval do def settings RailsSettings::ScopedSettings.for_thing(self) end scope :with_settings, :joins => "JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}')", :select => "DISTINCT #{self.table_name}.*" scope :with_settings_for, lambda { |var| { :joins => "JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}') AND settings.var = '#{var}'" } } scope :without_settings, :joins => "LEFT JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}')", :conditions => 'settings.id IS NULL' scope :without_settings_for, lambda { |var| { :joins => "LEFT JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}') AND settings.var = '#{var}'", :conditions => 'settings.id IS NULL' } } end end end end end end What I don't understand is why he uses class_eval on ActiveRecord::Base, wasn't it easier if he just open the ActiveRecord::Base class and define the functions? Specially that there's nothing dynamic in the block (What I mean by dynamic is when you do class_eval or instance_eval on a string containing variables) something like this: module ActiveRecord class Base def self.has_settings class_eval do def settings RailsSettings::ScopedSettings.for_thing(self) end scope :with_settings, :joins => "JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}')", :select => "DISTINCT #{self.table_name}.*" scope :with_settings_for, lambda { |var| { :joins => "JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}') AND settings.var = '#{var}'" } } scope :without_settings, :joins => "LEFT JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}')", :conditions => 'settings.id IS NULL' scope :without_settings_for, lambda { |var| { :joins => "LEFT JOIN settings ON (settings.thing_id = #{self.table_name}.#{self.primary_key} AND settings.thing_type = '#{self.base_class.name}') AND settings.var = '#{var}'", :conditions => 'settings.id IS NULL' } } end end end end I understand the second class_eval (before the def settings) is to define functions on the fly on every class that 'has_settings' right ? Same question here, I think he could use "def self.settings" instead of "class_eval.... def settings", no ?

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  • JQuery - Pass variables to PHP script via AJAX call and then display file

    - by hfidgen
    Hiya, I'm trying to generate Outlook event files for my events, doing so on the fly as and when someone requests it by pressing a link on the page. Here's what i've got so far, but I can't find out how to get the browser to download the content which is returned. I know how I could do this if I sent everything via _GET, but I'd prefer to do it via _POST, hence I'm going down this route.. Any thoughts? Thanks! HTML / Javascript <script> $(function() { $(".button").click(function() { // validate and process form // first hide any error messages var start = $("input#start").val(); var end = $("input#end").val(); var dataString = 'start='+ start + '&end=' + end; $.ajax({ type: "POST", url: "/calendar.php", data: dataString, success: function(data) { //Need to return the file contents somehow! } }); return false; }); }); </script> <form name="calendar" method="post" action=""> <input type="hidden" name="start" id="start" value="<?php echo $start; ?>" /> <input type="hidden" name="end" id="end" value="<?php echo $end; ?>" /> <input type="submit" name="submit" class="button" id="submit_btn" value="Outlook" /> </fieldset> </form> PHP File <?php if (isset($_POST['start'])) { $start = $_POST['start']; $end = $_POST['end']; $c = header("Content-Type: text/Calendar"); $c .= header("Content-Disposition: inline; filename=calendar.ics"); $c .= "BEGIN:VCALENDAR\n"; $c .= "VERSION:2.0\n"; $c .= "PRODID:-//xxxyyyy//NONSGML //EN\n"; $c .= "METHOD:REQUEST\n"; // requied by Outlook $c .= "BEGIN:VEVENT\n"; $c .= "UID:". $start . $end ."-" . "-xxxxyyyy.com\n"; // required by Outlook $c .= "DTSTAMP:".date('Ymd').'T'.date('His')."\n"; // required by Outlook $c .= "DTSTART:20080413T000000\n"; $c .= "SUMMARY:" . "\n"; $c .= "DESCRIPTION:" . "\n"; $c .= "END:VEVENT\n"; $c .= "END:VCALENDAR\n"; echo $c; } else { echo "Sorry you can't access this page directly"; } ?>

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  • How do you remind your Scrum Product Owner about his promises/actions?

    - by Felix Ogg
    ** EDIT: Rephrased the question to re-focus ** Our Scrum team meets as seldomly as possible, but we meet with the product owner every chance we get. We track everyone's agreed action points (particularly theirs). We are 100% agile, but our product owner lives in traditional world, we remain off-site. We facilitate him in crossing over to our fast-paced world. There's not much wrong. The team and the PO are in good spirits. PO is present at every meeting and positively energized. Just imagine this person as a 70 year old, slow grandpa, who is forgetful, yet kind. In reality he isn't, but he is used to a working environment (public servants) that is much slooooower. Manyana-manyana etc. It is frustrating for my team to cooperate: PO lives in a non-prioritized environment, and everyone in it has learned the productivity-technique of NGTD (Not Getting Things Done). He WANTS to, it's just that he forgets or 'sinks' somewhere along the away. We have experimented with a text file, maintained by the Scrum master (low-tech), which he broadcasts by e-mail every day JIRA, our issue tracker. Turns out this is nice for programmers, but too steep for 'regular people' I Googled for Issue tracking webtools but came up empty handed: All tools are aimed at IT issue tracking, instead of meeting action point tracking/planning for mere mortals. I did find TODO-lists like RememberTheMilk, but they don't track comments, and - to be honest - I doubt we could get our product owner to use it (too complicated). We have three requirements: Register action points, assign to a team member and a deadline Offer anyone to 'comment' on progress of any action point Do not build our own tool from scratch We do not need: - impressive authorization models, - multi-project, - workflow, - crosslinking. Is there any trick/tool you use to assist your product owner 'fly' like the rest of the rest of the team? Communication before tools I agree with the general consensus that one should not try to apply technology to a communication problem, however in this case I am merely looking for a tool to save me time in setting up prioritized lists. I found www.thymer.com today, may be what I am looking for. The guys are cool. It is getting rather feature-bloated though.

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  • Design patterns and interview question

    - by user160758
    When I was learning to code, I read up on the design patterns like a good boy. Long after this, I started to actually understand them. Design discussions such as those on this site constantly try to make the rules more and more general, which is good. But there is a line, over which it becomes over-analysis starts to feed off itself and as such I think begins to obfuscate the original point - for example the "What's Alternative to Singleton" post and the links contained therein. http://stackoverflow.com/questions/1300655/whats-alternative-to-singleton I say this having been asked in both interviews I’ve had over the last 2 weeks what a singleton is and what criticisms I have of it. I have used it a few times for items such as user data (simple key-value eg. last file opened by this user) and logging (very common i'm sure). I've never ever used it just to have what is essentially global application data, as this is clearly stupid. In the first interview, I reply that I have no criticisms of it. He seemed disappointed by this but as the job wasn’t really for me, I forgot about it. In the next one, I was asked again and, as I wanted this job, I thought about it on the spot and made some objections, similar to those contained in the post linked to above (I suggested use of a factory or dependency injection instead). He seemed happy with this. But my problem is that I have used the singleton without ever using it in this kind of stupid way, which I had to describe on the spot. Using it for global data and the like isn’t something I did then realised was stupid, or read was stupid so didn’t do, it was just something I knew was stupid from the start. Essentially I’m supposed to be able to think of ways of how to misuse a pattern in the interview? Which class of programmers can best answer this question? The best ones? The medium ones? I'm not sure.... And these were both bright guys. I read more than enough to get better at my job but had never actually bothered to seek out criticisms of the most simple of the design patterns like this one. Do people think such questions are valid and that I ought to know the objections off by heart? Or that it is reasonable to be able to work out what other people who are missing the point would do on the fly? Or do you think I’m at least partially right that the question is too unsubtle and that the questions ought to be better thought out in order to make sure only good candidates can answer. PS. Please don’t think I’m saying that I’m just so clever that I know everything automatically - I’ve learnt the hard way like everyone else. But avoiding global data is hardly revolutionary.

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  • Resize AIR app window while dragging

    - by matt lohkamp
    So I've noticed Windows 7 has a disturbing tendency to prevent you from dragging the title bar of windows off the top of the screen. If you try - in this case, using an air app with a draggable area at the bottom of the window, allowing you to push the top of the window up past the screen - it just kicks the window back down far enough that the title bar is at the top of what it considers the 'visible area.' One solution would be to resize the app window as it moves, so that the title bar is always where windows wants it. How would you resize the window while you're dragging it, though? Would you do it like this? dragHitArea.addEventListener(MouseEvent.MOUSE_DOWN, function(e:MouseEvent):void{ stage.nativeWindow.height += 50; stage.nativeWindow.startMove(); stage.nativeWindow.height -= 50; }); see what's going on there? When I click, I'm doing startMove(), which is hooking into the OS' function for dragging a window around. I'm also increasing and decreasing the height of the window by 50 pixels - which should give me no net increase, right? Wrong - the first '.height +=' gets executed, but the '.height -=' after the .startMove() never runs. Why? update - If you're curious, I'm programming an air widget with fly-out menus which expand rightwards and upwards - and since those element can only be displayed within the boundaries of the application window itself (even though the window is set to be chromeless and transparent) I have to expand the application's borders to include the area that the menu 'pops up' into. In the extreme case, with the widget positioned bottom left, and the menus expanded completely across to the right side and top edge of the screen, the application area could very well cover the entire desktop. The problem is, when it's expanded like this, if the user drags it up and to the right, it causes the 'title bar' area of the application window to move above the top edge of the desktop area, where it would normally be unreachable; and Windows automatically re-positions the window back below that edge once the .startMove() operation is completed. So what I want to do is continually resize the height of the application so that the visual effect will be the same for the user, but for the benefit of the operating system the window's title bar will never be above that top boundary of the desktop area.

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  • Combining FileStream and MemoryStream to avoid disk accesses/paging while receiving gigabytes of data?

    - by w128
    I'm receiving a file as a stream of byte[] data packets (total size isn't known in advance) that I need to store somewhere before processing it immediately after it's been received (I can't do the processing on the fly). Total received file size can vary from as small as 10 KB to over 4 GB. One option for storing the received data is to use a MemoryStream, i.e. a sequence of MemoryStream.Write(bufferReceived, 0, count) calls to store the received packets. This is very simple, but obviously will result in out of memory exception for large files. An alternative option is to use a FileStream, i.e. FileStream.Write(bufferReceived, 0, count). This way, no out of memory exceptions will occur, but what I'm unsure about is bad performance due to disk writes (which I don't want to occur as long as plenty of memory is still available) - I'd like to avoid disk access as much as possible, but I don't know of a way to control this. I did some testing and most of the time, there seems to be little performance difference between say 10 000 consecutive calls of MemoryStream.Write() vs FileStream.Write(), but a lot seems to depend on buffer size and the total amount of data in question (i.e the number of writes). Obviously, MemoryStream size reallocation is also a factor. Does it make sense to use a combination of MemoryStream and FileStream, i.e. write to memory stream by default, but once the total amount of data received is over e.g. 500 MB, write it to FileStream; then, read in chunks from both streams for processing the received data (first process 500 MB from the MemoryStream, dispose it, then read from FileStream)? Another solution is to use a custom memory stream implementation that doesn't require continuous address space for internal array allocation (i.e. a linked list of memory streams); this way, at least on 64-bit environments, out of memory exceptions should no longer be an issue. Con: extra work, more room for mistakes. So how do FileStream vs MemoryStream read/writes behave in terms of disk access and memory caching, i.e. data size/performance balance. I would expect that as long as enough RAM is available, FileStream would internally read/write from memory (cache) anyway, and virtual memory would take care of the rest. But I don't know how often FileStream will explicitly access a disk when being written to. Any help would be appreciated.

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  • getting Cannot identify image file when trying to create thumbnail in django

    - by Mo J. Mughrabi
    Am trying to create a thumbnail in django, am trying to build a custom class specifically to be used for generating thumbnails. As following from StringIO import StringIO from PIL import Image class Thumbnail(object): source = '' size = (50, 50) output = '' def __init__(self): pass @staticmethod def load(src): self = Thumbnail() self.source = src return self def generate(self, size=(50, 50)): if not isinstance(size, tuple): raise Exception('Thumbnail class: The size parameter must be an instance of a tuple.') self.size = size # resize properties box = self.size factor = 1 fit = True image = Image.open(self.source) # Convert to RGB if necessary if image.mode not in ('L', 'RGB'): image = image.convert('RGB') while image.size[0]/factor > 2*box[0] and image.size[1]*2/factor > 2*box[1]: factor *=2 if factor > 1: image.thumbnail((image.size[0]/factor, image.size[1]/factor), Image.NEAREST) #calculate the cropping box and get the cropped part if fit: x1 = y1 = 0 x2, y2 = image.size wRatio = 1.0 * x2/box[0] hRatio = 1.0 * y2/box[1] if hRatio > wRatio: y1 = int(y2/2-box[1]*wRatio/2) y2 = int(y2/2+box[1]*wRatio/2) else: x1 = int(x2/2-box[0]*hRatio/2) x2 = int(x2/2+box[0]*hRatio/2) image = image.crop((x1,y1,x2,y2)) #Resize the image with best quality algorithm ANTI-ALIAS image.thumbnail(box, Image.ANTIALIAS) # save image to memory temp_handle = StringIO() image.save(temp_handle, 'png') temp_handle.seek(0) self.output = temp_handle return self def get_output(self): return self.output.read() the purpose of the class is so i can use it inside different locations to generate thumbnails on the fly. The class works perfectly, I've tested it directly under a view.. I've implemented the thumbnail class inside the save method of the forms to resize the original images on saving. in my design, I have two fields for thumbnails. I was able to generate one thumbnail, if I try to generate two it crashes and I've been stuck for hours not sure whats the problem. Here is my model class Image(models.Model): article = models.ForeignKey(Article) title = models.CharField(max_length=100, null=True, blank=True) src = models.ImageField(upload_to='publication/image/') r128 = models.ImageField(upload_to='publication/image/128/', blank=True, null=True) r200 = models.ImageField(upload_to='publication/image/200/', blank=True, null=True) uploaded_at = models.DateTimeField(auto_now=True) Here is my forms class ImageForm(models.ModelForm): """ """ class Meta: model = Image fields = ('src',) def save(self, commit=True): instance = super(ImageForm, self).save(commit=True) file = Thumbnail.load(instance.src) instance.r128 = SimpleUploadedFile( instance.src.name, file.generate((128, 128)).get_output(), content_type='image/png' ) instance.r200 = SimpleUploadedFile( instance.src.name, file.generate((200, 200)).get_output(), content_type='image/png' ) if commit: instance.save() return instance the strange part is, when i remove the line which contains instance.r200 in the form save. It works fine, and it does the thumbnail and stores it successfully. Once I add the second thumbnail it fails.. Any ideas what am doing wrong here? Thanks Update: I tried earlier doing the following but I still got the same error class ImageForm(models.ModelForm): """ """ class Meta: model = Image fields = ('src',) def save(self, commit=True): instance = super(ImageForm, self).save(commit=True) instance.r128 = SimpleUploadedFile( instance.src.name, Thumbnail.load(instance.src).generate((128, 128)).get_output(), content_type='image/png' ) instance.r200 = SimpleUploadedFile( instance.src.name, Thumbnail.load(instance.src).generate((200, 200)).get_output(), content_type='image/png' ) if commit: instance.save() return instance

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  • Can I use the [] operator in C++ to create virtual arrays

    - by Shane MacLaughlin
    I have a large code base, originally C ported to C++ many years ago, that is operating on a number of large arrays of spatial data. These arrays contain structs representing point and triangle entities that represent surface models. I need to refactor the code such that the specific way these entities are stored internally varies for specific scenarios. For example if the points lie on a regular flat grid, I don't need to store the X and Y coordinates, as they can be calculated on the fly, as can the triangles. Similarly, I want to take advantage of out of core tools such as STXXL for storage. The simplest way of doing this is replacing array access with put and get type functions, e.g. point[i].x = XV; becomes Point p = GetPoint(i); p.x = XV; PutPoint(i,p); As you can imagine, this is a very tedious refactor on a large code base, prone to all sorts of errors en route. What I'd like to do is write a class that mimics the array by overloading the [] operator. As the arrays already live on the heap, and move around with reallocs, the code already assumes that references into the array such as point *p = point + i; may not be used. Is this class feasible to write? For example writing the methods below in terms of the [] operator; void MyClass::PutPoint(int Index, Point p) { if (m_StorageStrategy == RegularGrid) { int xoffs,yoffs; ComputeGridFromIndex(Index,xoffs,yoffs); StoreGridPoint(xoffs,yoffs,p.z); } else m_PointArray[Index] = p; } } Point MyClass::GetPoint(int Index) { if (m_StorageStrategy == RegularGrid) { int xoffs,yoffs; ComputeGridFromIndex(Index,xoffs,yoffs); return GetGridPoint(xoffs,yoffs); // GetGridPoint returns Point } else return m_PointArray[Index]; } } My concern is that all the array classes I've seen tend to pass by reference, whereas I think I'll have to pass structs by value. I think it should work put other than performance, can anyone see any major pitfalls with this approach. n.b. the reason I have to pass by value is to get point[a].z = point[b].z + point[c].z to work correctly where the underlying storage type varies.

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  • Can I have a workspace that is both a git workspace and a svn workspace?

    - by Troy
    I have checked out now a local working copy of a codebase that lives in an svn repo. It's a big Java project that I use Eclipse to develop in. Eclipse of course builds everything on the fly, in it's own way with all the binaries ending up in [project root]/bin. That's perfectly fine with me, for development, but when the build runs on the build server, it looks quite a lot different (maven build, binaries end up in a different directory structure, etc). Sometimes I need to recreate the build server environment on my local development system to debug the build or what have you, so I usually end up downloading an entirely new working copy into a new workspace and running the build from there (prevents cluttering my development workspace with all the build artifacts and dirtying up the working copy). Of course sometimes I'm interested in running the full build on code that I don't want to check in yet, so I will manually copy over the "development" workspace onto the "build" workspace. Besides taking a lot of extra time copying a lot of files that I don't actually need (just overlaying the new over the old), this also screws up my svn metadata, meaning that I can't check in changes from that "build workspace" working copy, and I often end up having to re-download the code to get it back into a known state. So I'm thinking I make my svn working copy a local git repo, then "check out" the in-development code from the svn working copy/git master, into the local build workspace. Then I can build, revert my changes, have all the advantages of a version controlled working copy in the build workspace. Then if I need to make changes to the build, push those back into the git master (which is also a svn working copy), then check them into the main svn repo. |-------------| |main svn repo| <------- |---------------------| |-------------| |svn working copy | <------- |--------------------| | (svn dev workspace/ | | non-svn-versioned | | git master) | | build workspace | |---------------------| | (git working copy) | |--------------------| Just switching everything to git would obviously be better, but, big company, too many people using svn, too costly to change everything, etc. We're stuck with svn as the main repo for now. BTW, I know there is a maven plugin for Eclipse and everything, I'm mainly interested to know if there is a way to maintain a workspace that is both a git working copy and an svn working copy. Actually any distributed version control system would probably work (hg possibly?). Advice? How does everybody else handle this situation of having a to manage both a "development" build process and a "production" build process?

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  • DataGridView live display of datatable using virtual mode

    - by Chris
    I have a DataGridView that will display records (log entries) from a database. The amount of records that can exist at a time is very large. I would like to use the virtual mode feature of the DataGridView to display a page of data, and to minimize the amount of data that has to be transferred across a network at a given time. Polling for data is out of the question. There will be several clients running at a time, all of which are on the same network and viewing the records. If they all poll for data, the network will run very slowly. The data is read-only to the user; they won't be able to edit any of it, just view it. I need to know when updates occur in the database, and I need to update the screen with those updates accordingly using virtual mode. If a page of data a user is viewing contains data that has change, he/she will see those updates on that page. If updates were made to data in the database, but not in the data the user is viewing, then not much really changes on the user screen (Maybe just the scroll bar if records were added or removed). My current approach is using SQL server change tracking with the sync framework. Each client has a local SQL Server CE instance and database file that is kept in sync with the main database server. I use the information from the synchronization event to see if any changes were made to the main database and were sync'ed to the client. I need to use the DataGridView virtual mode here because I can't have thousands of records loaded into the DataGridView at once, otherwise memory usage goes through the roof. The main challenge right now is knowing how to use virtual mode to provide a seamless experience to the user by allowing them to scroll up and down through the records, and also have records update on the fly without interfering with the user inappropriately. Has anybody dealt with this issue before, and if so, where I can see how they did it? I've gone through some of the MSDN documentation and examples on virtual mode. So far, I haven't found documentation and/or examples on their site that explains how to do what I am trying to accomplish.

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  • Table row height in Internet Explorer

    - by Fritz H
    I have the following table: <table> <tr> <td style="height: 7px; width: 7px"> A1 </td> <td style="height: 7px"> B1 </td> <td style="height: 7px; width: 7px"> C1 </td> </tr> <tr> <td style="width: 7px"> A2 </td> <td> B2 </td> <td style="width: 7px"> C2 </td> </tr> <tr> <td style="height: 7px; width: 7px"> A3 </td> <td style="height: 7px"> B3 </td> <td style="height: 7px; width: 7px"> C3 </td> </tr> </table> The basic idea is that the first row must be 7 pixels high. The left- and rightmost cells (A1 and C1) must be 7px wide, and the middle cell (B1) must scale according to the width of the table. The same goes for the bottom row (A3, B3, C3). The middle row, however, needs to scale in height - in other words, it needs to be (tableheight - 14px). The left- and rightmost cells (A2, C2) need to be 7 pixels wide. An example: 7px x 7px |------|-------------------------|------| --- +------+-------------------------+------+ | | | | | | 7px | | | | | | | | | --- +------+-------------------------+------+ | | | | | | | | | | | | | | | | | | | | | y | | | | | | | | | | | | | | | | | | | | | | | | --- +------+-------------------------+------+ | | | | | | 7px | | | | | | | | | --- +------+-------------------------+------+ HOWEVER: In Internet Explorer, the widths work fine (columns A and C are 7px, column B scales dynamically) - but the heights don't. Rows 1, 2 and 3 turn out to be exactly 33% of the height of the table, no matter what I do. Unfortunately I have to use this table, so replacing it with a set of DIVs is not an option. I have the following DOCTYPE: <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> I need to keep this, as some other elements on the page rely on some complex CSS-based layouts. Can anyone point me in the right direction to whip this into shape for IE? EDIT: Should have mentioned earlier - this table is resized on the fly using javascript.

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  • IPhone iOs 5, need help getting my tab bar at the top to work

    - by Patrick
    I wanted to have the tab bar at the top. So i created a new project in XCode. Added a view and then inside that view i added (scrollbar, text and another view). See picture. What i wanted was to have my tab bar at the top. Then in the middle would be the contents from the tab bar and below it a small copyright text. See picture. No idea how to make this correctly. I have tried to create the UITabBarController on the fly and then assign it into the view at the top. (Top white space on the picture dedicated for the tab bar). Here is my code to init the MainWindow. MainWindow.h #import <UIKit/UIKit.h> @class Intro; @interface MainWindow : UIViewController @property (strong, nonatomic) IBOutlet UIScrollView *mainContentFrame; @property (strong, nonatomic) IBOutlet UIView *mainTabBarView; @property (strong, nonatomic) UITabBarController *mainTabBar; @property (nonatomic, strong) Intro *intro; // Trying to get this tab to show in the tab bar @end MainWindow.m #import "MainWindow.h" #import "Intro.h" @interface MainWindow () @end @implementation MainWindow @synthesize mainContentFrame = _mainContentFrame; @synthesize mainTabBarView = _mainTabBarView; @synthesize mainTabBar = _mainTabBar; @synthesize intro = _intro; - (id)initWithNibName:(NSString *)nibNameOrNil bundle:(NSBundle *)nibBundleOrNil { self = [super initWithNibName:nibNameOrNil bundle:nibBundleOrNil]; if (self) { // Custom initialization } return self; } - (void)viewDidLoad { _intro = [[Intro alloc] init]; NSArray *allViews = [[NSArray alloc] initWithObjects:_intro, nil]; [super viewDidLoad]; // Do any additional setup after loading the view. _mainTabBar = [[UITabBarController alloc] init]; [_mainTabBar setViewControllers:allViews]; [_mainTabBarView.window addSubview:_mainTabBar.tabBarController.view]; } - (void)viewDidUnload { [self setMainTabBar:nil]; [self setMainContentFrame:nil]; [self setMainContentFrame:nil]; [super viewDidUnload]; // Release any retained subviews of the main view. } - (BOOL)shouldAutorotateToInterfaceOrientation:(UIInterfaceOrientation)interfaceOrientation { return (interfaceOrientation == UIInterfaceOrientationPortrait); } @end What am i missing to get this to work? Wanted the content to end up in the scrollview so that all tabs are scrollable.

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  • jQuery .load(), don't show new content until images loaded

    - by Jarred
    Hi. I have been working on a jQuery photo slideshow. It scales the images to the browser size, and slides them left and right. There is no pre-determined size or aspect ratio, the script does everything on the fly. It requires that all images be fully loaded, so it can custom resize each individual image based on it's own aspect ratio ( width():height(), etc ), calculate the width of containing div, and calculate the slide distance from one image to another. As a stand-alone, it works pretty well (despite my lack of skills)! I simply hide the slideshow containing div at (document).ready, allow the images to load, then run the slideshow prep scripts at (window).load. Once this is done, it only then makes the slideshow divs, images, etc appear, properly sized, positioned and ready to roll. The ultimate goal is to be able to load in any number of slideshows without refreshing the page. The point of this is to be able to play uninterrupted background music. I know music on websites is annoying, but the target market likes it, a lot! I am using (target).load(page.php .element, function prepInsertNewShow() { //Prepare slideshow resizeImages(); slideArray(); //Show slideshow (target).fadeIn(); }); and it definitely works! The problem is that I cannot find a way to hold off on preparing and showing the new content until the images have finished loading. It is running the slideshow prep scripts (which are totally dependent on the images being fully loaded), before the images are loaded. This results in a completely jacked up show! What I want to do is this - (target).load(page.php .element, function prepInsertNewShow() { //Wait until images are loaded $('img').load( function() { //Prepare slideshow resizeImages(); slideArray(); //Show slideshow (target).fadeIn(); } }); But this doesn't seem to work, the new content is never shown. You can see a live version here. The initial gallery loads correctly, everything looks good. The only nav link that works is Galleries Engagement, which will load a new show (a containing div with multiple <img> tags). You will see that the images are not centered, the containing div and slide distances are much too small, as they were calculated using images that were not actually loaded. Is there any way I can delay handling and showing new content until it is fully loaded? Any suggestions would be most appreciated, thanks for your time! PS - It just occurred to me while typing this that a decent solution may be to insert "width=x" height="x" into the <img> tags, so the script can work from those values, even if the images have not loaded...hmm...

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  • AudioTrack lag: obtainBuffer timed out

    - by BTR
    I'm playing WAVs on my Android phone by loading the file and feeding the bytes into AudioTrack.write() via the FileInputStream BufferedInputStream DataInputStream method. The audio plays fine and when it is, I can easily adjust sample rate, volume, etc on the fly with nice performance. However, it's taking about two full seconds for a track to start playing. I know AudioTrack has an inescapable delay, but this is ridiculous. Every time I play a track, I get this: 03-13 14:55:57.100: WARN/AudioTrack(3454): obtainBuffer timed out (is the CPU pegged?) 0x2e9348 user=00000960, server=00000000 03-13 14:55:57.340: WARN/AudioFlinger(72): write blocked for 233 msecs, 9 delayed writes, thread 0xba28 I've noticed that the delayed write count increases by one every time I play a track -- even across multiple sessions -- from the time the phone has been turned on. The block time is always 230 - 240ms, which makes sense considering a minimum buffer size of 9600 on this device (9600 / 44100). I've seen this message in countless searches on the Internet, but it usually seems to be related to not playing audio at all or skipping audio. In my case, it's just a delayed start. I'm running all my code in a high priority thread. Here's a truncated-yet-functional version of what I'm doing. This is the thread callback in my playback class. Again, this works (only playing 16-bit, 44.1kHz, stereo files right now), it just takes forever to start and has that obtainBuffer/delayed write message every time. public void run() { // Load file FileInputStream mFileInputStream; try { // mFile is instance of custom file class -- this is correct, // so don't sweat this line mFileInputStream = new FileInputStream(mFile.path()); } catch (FileNotFoundException e) {} BufferedInputStream mBufferedInputStream = new BufferedInputStream(mFileInputStream, mBufferLength); DataInputStream mDataInputStream = new DataInputStream(mBufferedInputStream); // Skip header try { if (mDataInputStream.available() > 44) mDataInputStream.skipBytes(44); } catch (IOException e) {} // Initialize device mAudioTrack = new AudioTrack(AudioManager.STREAM_MUSIC, ConfigManager.SAMPLE_RATE, AudioFormat.CHANNEL_CONFIGURATION_STEREO, AudioFormat.ENCODING_PCM_16BIT, ConfigManager.AUDIO_BUFFER_LENGTH, AudioTrack.MODE_STREAM); mAudioTrack.play(); // Initialize buffer byte[] mByteArray = new byte[mBufferLength]; int mBytesToWrite = 0; int mBytesWritten = 0; // Loop to keep thread running while (mRun) { // This flag is turned on when the user presses "play" while (mPlaying) { try { // Check if data is available if (mDataInputStream.available() > 0) { // Read data from file and write to audio device mBytesToWrite = mDataInputStream.read(mByteArray, 0, mBufferLength); mBytesWritten += mAudioTrack.write(mByteArray, 0, mBytesToWrite); } } catch (IOException e) { } } } } If I can get past the artificially long lag, I can easily deal with the inherit latency by starting my write at a later, predictable position (ie, skip past the minimum buffer length when I start playing a file).

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  • With custom binding for Masonry how would I resize the container appropriately

    - by BigDubb
    Based on the suggestion give here, and the information given here on how to make a custom bindingHandler for a forEach, I decided to attempt to write my own custom binding for a forEach and Masonry. Because the elements are added on the fly the redrawing and moving around of elements to fill the space doesn't occur. So, this functionality needed to be moved after the elements have been rendered or called after each item has been added. Here is my bindingHandler ko.bindingHandlers.masonry = { init: function (element, valueAccessor, allBindingsAccessor) { var $element = $(element), originalContent = $element.html(); $element.data("original-content", originalContent); //var msnry = new Masonry($element); return { controlsDescendantBindings: true } }, update: function (element, valueAccessor, allBindingsAccessor) { var value = ko.utils.unwrapObservable(valueAccessor()), //get the list of items items = value.items(), //get a jQuery reference to the element $element = $(element), //get the current content of the element elementContent = $element.data("original-content"); $element.html(""); var container = $element[0]; var msnry = new Masonry(container); for (var index = 0; index < items.length; index++) { (function () { //get the list of items var item = ko.utils.unwrapObservable(items[index]), $childElement = $(elementContent); ko.applyBindings(item, $childElement[0]); //add the child to the parent $element.append($childElement); msnry.appended($childElement[0]); })(); msnry.layout(); msnry.bindResize(); } } }; and the HTML implementing the handler. <div id="criteriaContainer" data-bind="masonry: { items: SearchItems.Items }"> <div class="searchCriterion control-group"> <label class="control-label" data-bind="text: Description"></label> <div class="controls"> <input type="hidden" data-bind="value: Value, select2: { minimumInputLength: 3, queryUri: SearchUri(), placeholder: Placeholder(), allowClear: true }" style="width: 450px"> </div> <p data-bind="text: Value"></p> </div> </div> When this shows up on the page It stacks all if the elements rendered via the append method right on top of each other. You can see in my bindingHandler I am calling bindResize as well as layout(), neither of which seem to be having any effect. Here's a screenshot of what it looks like in the UI.

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  • C# 4: The Curious ConcurrentDictionary

    - by James Michael Hare
    In my previous post (here) I did a comparison of the new ConcurrentQueue versus the old standard of a System.Collections.Generic Queue with simple locking.  The results were exactly what I would have hoped, that the ConcurrentQueue was faster with multi-threading for most all situations.  In addition, concurrent collections have the added benefit that you can enumerate them even if they're being modified. So I set out to see what the improvements would be for the ConcurrentDictionary, would it have the same performance benefits as the ConcurrentQueue did?  Well, after running some tests and multiple tweaks and tunes, I have good and bad news. But first, let's look at the tests.  Obviously there's many things we can do with a dictionary.  One of the most notable uses, of course, in a multi-threaded environment is for a small, local in-memory cache.  So I set about to do a very simple simulation of a cache where I would create a test class that I'll just call an Accessor.  This accessor will attempt to look up a key in the dictionary, and if the key exists, it stops (i.e. a cache "hit").  However, if the lookup fails, it will then try to add the key and value to the dictionary (i.e. a cache "miss").  So here's the Accessor that will run the tests: 1: internal class Accessor 2: { 3: public int Hits { get; set; } 4: public int Misses { get; set; } 5: public Func<int, string> GetDelegate { get; set; } 6: public Action<int, string> AddDelegate { get; set; } 7: public int Iterations { get; set; } 8: public int MaxRange { get; set; } 9: public int Seed { get; set; } 10:  11: public void Access() 12: { 13: var randomGenerator = new Random(Seed); 14:  15: for (int i=0; i<Iterations; i++) 16: { 17: // give a wide spread so will have some duplicates and some unique 18: var target = randomGenerator.Next(1, MaxRange); 19:  20: // attempt to grab the item from the cache 21: var result = GetDelegate(target); 22:  23: // if the item doesn't exist, add it 24: if(result == null) 25: { 26: AddDelegate(target, target.ToString()); 27: Misses++; 28: } 29: else 30: { 31: Hits++; 32: } 33: } 34: } 35: } Note that so I could test different implementations, I defined a GetDelegate and AddDelegate that will call the appropriate dictionary methods to add or retrieve items in the cache using various techniques. So let's examine the three techniques I decided to test: Dictionary with mutex - Just your standard generic Dictionary with a simple lock construct on an internal object. Dictionary with ReaderWriterLockSlim - Same Dictionary, but now using a lock designed to let multiple readers access simultaneously and then locked when a writer needs access. ConcurrentDictionary - The new ConcurrentDictionary from System.Collections.Concurrent that is supposed to be optimized to allow multiple threads to access safely. So the approach to each of these is also fairly straight-forward.  Let's look at the GetDelegate and AddDelegate implementations for the Dictionary with mutex lock: 1: var addDelegate = (key,val) => 2: { 3: lock (_mutex) 4: { 5: _dictionary[key] = val; 6: } 7: }; 8: var getDelegate = (key) => 9: { 10: lock (_mutex) 11: { 12: string val; 13: return _dictionary.TryGetValue(key, out val) ? val : null; 14: } 15: }; Nothing new or fancy here, just your basic lock on a private object and then query/insert into the Dictionary. Now, for the Dictionary with ReadWriteLockSlim it's a little more complex: 1: var addDelegate = (key,val) => 2: { 3: _readerWriterLock.EnterWriteLock(); 4: _dictionary[key] = val; 5: _readerWriterLock.ExitWriteLock(); 6: }; 7: var getDelegate = (key) => 8: { 9: string val; 10: _readerWriterLock.EnterReadLock(); 11: if(!_dictionary.TryGetValue(key, out val)) 12: { 13: val = null; 14: } 15: _readerWriterLock.ExitReadLock(); 16: return val; 17: }; And finally, the ConcurrentDictionary, which since it does all it's own concurrency control, is remarkably elegant and simple: 1: var addDelegate = (key,val) => 2: { 3: _concurrentDictionary[key] = val; 4: }; 5: var getDelegate = (key) => 6: { 7: string s; 8: return _concurrentDictionary.TryGetValue(key, out s) ? s : null; 9: };                    Then, I set up a test harness that would simply ask the user for the number of concurrent Accessors to attempt to Access the cache (as specified in Accessor.Access() above) and then let them fly and see how long it took them all to complete.  Each of these tests was run with 10,000,000 cache accesses divided among the available Accessor instances.  All times are in milliseconds. 1: Dictionary with Mutex Locking 2: --------------------------------------------------- 3: Accessors Mostly Misses Mostly Hits 4: 1 7916 3285 5: 10 8293 3481 6: 100 8799 3532 7: 1000 8815 3584 8:  9:  10: Dictionary with ReaderWriterLockSlim Locking 11: --------------------------------------------------- 12: Accessors Mostly Misses Mostly Hits 13: 1 8445 3624 14: 10 11002 4119 15: 100 11076 3992 16: 1000 14794 4861 17:  18:  19: Concurrent Dictionary 20: --------------------------------------------------- 21: Accessors Mostly Misses Mostly Hits 22: 1 17443 3726 23: 10 14181 1897 24: 100 15141 1994 25: 1000 17209 2128 The first test I did across the board is the Mostly Misses category.  The mostly misses (more adds because data requested was not in the dictionary) shows an interesting trend.  In both cases the Dictionary with the simple mutex lock is much faster, and the ConcurrentDictionary is the slowest solution.  But this got me thinking, and a little research seemed to confirm it, maybe the ConcurrentDictionary is more optimized to concurrent "gets" than "adds".  So since the ratio of misses to hits were 2 to 1, I decided to reverse that and see the results. So I tweaked the data so that the number of keys were much smaller than the number of iterations to give me about a 2 to 1 ration of hits to misses (twice as likely to already find the item in the cache than to need to add it).  And yes, indeed here we see that the ConcurrentDictionary is indeed faster than the standard Dictionary here.  I have a strong feeling that as the ration of hits-to-misses gets higher and higher these number gets even better as well.  This makes sense since the ConcurrentDictionary is read-optimized. Also note that I tried the tests with capacity and concurrency hints on the ConcurrentDictionary but saw very little improvement, I think this is largely because on the 10,000,000 hit test it quickly ramped up to the correct capacity and concurrency and thus the impact was limited to the first few milliseconds of the run. So what does this tell us?  Well, as in all things, ConcurrentDictionary is not a panacea.  It won't solve all your woes and it shouldn't be the only Dictionary you ever use.  So when should we use each? Use System.Collections.Generic.Dictionary when: You need a single-threaded Dictionary (no locking needed). You need a multi-threaded Dictionary that is loaded only once at creation and never modified (no locking needed). You need a multi-threaded Dictionary to store items where writes are far more prevalent than reads (locking needed). And use System.Collections.Concurrent.ConcurrentDictionary when: You need a multi-threaded Dictionary where the writes are far more prevalent than reads. You need to be able to iterate over the collection without locking it even if its being modified. Both Dictionaries have their strong suits, I have a feeling this is just one where you need to know from design what you hope to use it for and make your decision based on that criteria.

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  • Stream Music and Video Over the Internet with Windows Media Player 12

    - by DigitalGeekery
    A new feature in Windows Media Player 12, which is included with Windows 7, is being able to stream media over the web to other Windows 7 computers.  Today we will take a look at how to set it up and what you need to begin. Note: You will need to perform this process on each computer that you want to use. What You’ll Need Two computers running Windows 7 Home Premium, Professional, or Ultimate. The host, or home computer that you will be streaming the media from, cannot be on a public network or part of domain. Windows Live ID UPnP or Port Forwarding enabled on your home router Media files added to your Windows Media Player library Windows Live ID Sign up online for a Windows Live ID if you do not already have one. See the link below for a link to Windows Live.   Configuring the Windows 7 Computers Open Windows Media Player and go to the library section. Click on Stream and then “Allow Internet access to home media.”   The Internet Home Media Access pop up window will prompt you to link your Windows Live ID to a user account. Click “Link an online ID.” If you haven’t already installed the Windows Live ID Sign-In Assistant, you will be taken to Microsoft’s website and prompted to download it. Once you have completed the Windows Live download assistant install, you will see Windows Live ID online provider appear in the “Link Online IDs” window. Click on “Link Online ID.” Next, you’ll be prompted for a Windows Live ID and password. Enter your Windows Live ID and password and click “Sign In.” A pop up window will notify you that you have successfully allowed Internet access to home media. Now, you will have to repeat the exact same configuration on the 2nd Windows 7 computer. Once you have completed the same configuration on your 2nd computer, you might also need to configure your home router for port forwarding. If your router supports UPnP, you may not need to manually forward any ports on your router. So, this would be a good time to test your connection. Go to a nearby hotspot, or perhaps a neighbor’s house, and test to see if you can stream your media. If not, you’ll need to manually forward the ports. You can always choose to forward the ports anyway, just in case. Note: We tested on a Linksys WRT54GL router, which supports UPnP, and found we still needed to manually forward the ports. Finding the ports to forward on the router Open Windows Media Player and make sure you are in Library view. Click on “Stream” on the top menu, and select “Allow Internet access to home media.”   On the “Internet Home Media Access” window, click on “Diagnose connections.” The “Internet Streaming Diagnostic Tool” will pop up. Click on “Port forwarding information” near the bottom.   On the “Port Forwarding Information” window you will find both the Internal and External Port numbers you will need to forward on your router. The Internal port number should always be 10245. The external number will be different depending on your computer. Microsoft also recommends forwarding port 443. Configuring the Router Next, you’ll need to configure Port Forwarding on your home router. We will show you the steps for a Linksys WRT54GL router, however, the steps for port forwarding will vary from router to router. On the Linksys configuration page, click on the Administration Tab along the top, click the “Applications & Gaming Tab, and then the “Port Range Forward” tab below it. Under “Application,” type in a name. It can be any name you choose. In both the “Start” and “End” boxes, type the port number. Enter the IP address of your home computer in the IP address column. Click the check box under “Enable.” Do this for both the internal and external port numbers and port 443. When finished, click the “Save Settings” button. Note: It’s highly recommended that you configure your home computer with a static IP address When you’re ready to play your media over the Internet, open up Windows Media Player and look for your host computer and username listed under “Other Libraries.” Click on it expand the list to see your media libraries. Choose a library and a file to play. Now you can enjoy your streaming media over the Internet. Conclusion We found media streaming over the Internet to work fairly well. However, we did see a loss of quality with streaming video. Also, Recorded TV .wtv and dvr-ms files did not play at all. Check out our previous article to see how to stream media share and stream media between Windows 7 computers on your home network. Similar Articles Productive Geek Tips Enable Media Streaming in Windows Home Server to Windows Media PlayerFixing When Windows Media Player Library Won’t Let You Add FilesShare Digital Media With Other Computers on a Home Network with Windows 7Share and Stream Digital Media Between Windows 7 Machines On Your Home NetworkLearning Windows 7: Manage Your Music with Windows Media Player TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 PCmover Professional Stormpulse provides slick, real time weather data Geek Parents – Did you try Parental Controls in Windows 7? 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  • Customize your icons in Windows 7 and Vista

    - by Matthew Guay
    Want to change out the icons on your desktop and more?  Personalizing your icons is a great way to make your PC uniquely yours,, and today we show you how to grab unique icons, and default Winnows. to be your own. Change the icon for Computer, Recycle Bin, Network, and your User folder Right-click on the desktop, and select Personalize. Now, click the “Change desktop icons” link on the left sidebar in the Personalization window. The window looks slightly different in Windows Vista, but the link is the same. Select the icon you wish to change, and click the Change Icon button.  In Windows 7, you will also notice a box to choose whether or not to allow themes to change icons, and you can uncheck it if you don’t want themes to change your icon settings. You can select one of the other included icons, or click browse to find the icon you want.  Click Ok when you are finished. Change Folder icons You can easily change the icon on most folders in Windows Vista and 7.  Simply right-click on the folder and select properties. Click the Customize tab, and then click the Change Icon button.  This will open the standard dialog to change your icon, so proceed as normal. This basically just creates a hidden desktop.ini file in the folder containing the following or similar data: [.ShellClassInfo]IconFile=%SystemRoot%\system32\SHELL32.dllIconIndex=20 You could manually create or edit the file if you choose, instead of using the dialogs. Simply create a new text file named desktop.ini with this same information, or edit the existing one.  Change the IconFile line to the location of your icon. If you are pointing to a .ico file you should change the IconIndex line to 0 instead. Note that this isn’t available for all folders, for instance you can’t use this to change the icon for the Windows folder.   In Windows 7, please note that you cannot change the icon of folder inside a library.  So if you are browsing your Documents library and would like to change an icon in that folder, right-click on it and select Open folder location.  Now you can change the icon as above. And if you would like to change a Library’s icon itself, then check out this tutorial: Change Your Windows 7 Library Icons the Easy Way Change the icon of any file type Want to make you files easier to tell apart?  Check out our tutorial on how to simply do this: Change a File Type’s Icon in Windows 7 Change the icon of any Application Shortcut To change the icon of a shortcut on your desktop, start menu, or in Explorer, simply right-click on the icon and select Properties. In the Shortcut tab, click the Change Icon button. Now choose one of the other available icons or click browse to find the icon you want. Change Icons of Running Programs in the Windows 7 taskbar If your computer is running Windows 7, you can customize the icon of any program running in the taskbar!  This only works on applications that are running but not pinned to the taskbar, so if you want to customize a pinned icon you may want to unpin it before customizing it.  But the interesting thing about this trick is that it can customize any icon anything running in the taskbar, including things like Control Panel! Right-click or click and push up to open the jumplist on the icon, and then right-click on the program’s name and select Properties.  Here we are customizing Control Panel, but you can do this on any application icon. Now, click Change Icon as usual. Select an icon you want (We switched the Control Panel icon to the Security Shield), or click Browse to find another icon.  Click Ok when finished, and then close the application window. The next time you open the program (or Control Panel in our example), you will notice your new icon on its taskbar icon. Please note that this only works on applications that are currently running and are not pinned to the taskbar.  Strangely, if the application is pinned to the taskbar, you can still click Properties and change the icon, but the change will not show up. Change the icon on any Drive on your Computer You can easily change the icon on your internal hard drives and portable drives with the free Drive Icon Changer application.  Simply download and unzip the file (link below), and then run the application as administrator by right-clicking on the icon and selecting “Run as administrator”. Now, select the drive that you want to change the icon of, and select your desired icon file. Click Save, and Drive Icon Changer will let you know that the icon has been changed successfully. You will then need to reboot your computer to complete the changes.  Simply click Yes to reboot. Now, our Drive icon is changed from this default image: to a Laptop icon we chose! You can do this to any drive in your computer, or to removable drives such as USB flash drives.  When you change these drives icons, the new icon will appear on any computer you insert the drive into.  Also, if you wish to remove the icon change, simply run the Drive Icon Changer again and remove the icon path. Download Drive Icon Changer This application actually simply creates or edits a hidden Autorun.inf file on the top of your drive.  You can edit or create the file yourself by hand if you’d like; simply include the following information in the file, and save it in the top directory of your drive: [autorun]ICON=[path of your icon] Remove Arrow from shortcut icons Many people don’t like the arrow on the shortcut icon, and there are two easy ways to do this. If you’re running the 32 bit version of Windows Vista or 7, simply use the Vista Shortcut Overlay Remover. If your computer is running the 64 bit version of Windows Vista or 7, use the Ultimate Windows Tweaker instead.  Simply select the Additional Tweaks section, and check the “Remove arrows from Shortcut Icons.” For more info and download links check out this article: Disable Shortcut Icon Arrow Overlay in Windows 7 or Vista Closing: This gives you a lot of ways to customize almost any icon on your computer, so you can make it look just like you want it to.  Stay tuned for more great desktop customization articles from How-to Geek! Similar Articles Productive Geek Tips Change Start Menu to Use Small Icons in Windows 7 or VistaResize Icons Quickly in Windows 7 or Vista ExplorerRoundup: 16 Tweaks to Windows Vista Look & FeelRestore Missing Desktop Icons in Windows 7 or VistaClean Up Past Notification Icons in Windows Vista TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 PCmover Professional Change DNS servers on the fly with DNS Jumper Live PDF Searches PDF Files and Ebooks Converting Mp4 to Mp3 Easily Use Quick Translator to Translate Text in 50 Languages (Firefox) Get Better Windows Search With UltraSearch Scan News With NY Times Article Skimmer

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  • Recover Deleted Files on an NTFS Hard Drive from a Ubuntu Live CD

    - by Trevor Bekolay
    Accidentally deleting a file is a terrible feeling. Not being able to boot into Windows and undelete that file makes that even worse. Fortunately, you can recover deleted files on NTFS hard drives from an Ubuntu Live CD. To show this process, we created four files on the desktop of a Windows XP machine, and then deleted them. We then booted up the same machine with the bootable Ubuntu 9.10 USB Flash Drive that we created last week. Once Ubuntu 9.10 boots up, open a terminal by clicking Applications in the top left of the screen, and then selecting Accessories > Terminal. To undelete our files, we first need to identify the hard drive that we want to undelete from. In the terminal window, type in: sudo fdisk –l and press enter. What you’re looking for is a line that ends with HPSF/NTFS (under the heading System). In our case, the device is “/dev/sda1”. This may be slightly different for you, but it will still begin with /dev/. Note this device name. If you have more than one hard drive partition formatted as NTFS, then you may be able to identify the correct partition by the size. If you look at the second line of text in the screenshot above, it reads “Disk /dev/sda: 136.4 GB, …” This means that the hard drive that Ubuntu has named /dev/sda is 136.4 GB large. If your hard drives are of different size, then this information can help you track down the right device name to use. Alternatively, you can just try them all, though this can be time consuming for large hard drives. Now that you know the name Ubuntu has assigned to your hard drive, we’ll scan it to see what files we can uncover. In the terminal window, type: sudo ntfsundelete <HD name> and hit enter. In our case, the command is: sudo ntfsundelete /dev/sda1 The names of files that can recovered show up in the far right column. The percentage in the third column tells us how much of that file can be recovered. Three of the four files that we originally deleted are showing up in this list, even though we shut down the computer right after deleting the four files – so even in ideal cases, your files may not be recoverable. Nevertheless, we have three files that we can recover – two JPGs and an MPG. Note: ntfsundelete is immediately available in the Ubuntu 9.10 Live CD. If you are in a different version of Ubuntu, or for some other reason get an error when trying to use ntfsundelete, you can install it by entering “sudo apt-get install ntfsprogs” in a terminal window. To quickly recover the two JPGs, we will use the * wildcard to recover all of the files that end with .jpg. In the terminal window, enter sudo ntfsundelete <HD name> –u –m *.jpg which is, in our case, sudo ntfsundelete /dev/sda1 –u –m *.jpg The two files are recovered from the NTFS hard drive and saved in the current working directory of the terminal. By default, this is the home directory of the current user, though we are working in the Desktop folder. Note that the ntfsundelete program does not make any changes to the original NTFS hard drive. If you want to take those files and put them back in the NTFS hard drive, you will have to move them there after they are undeleted with ntfsundelete. Of course, you can also put them on your flash drive or open Firefox and email them to yourself – the sky’s the limit! We have one more file to undelete – our MPG. Note the first column on the far left. It contains a number, its Inode. Think of this as the file’s unique identifier. Note this number. To undelete a file by its Inode, enter the following in the terminal: sudo ntfsundelete <HD name> –u –i <Inode> In our case, this is: sudo ntfsundelete /dev/sda1 –u –i 14159 This recovers the file, along with an identifier that we don’t really care about. All three of our recoverable files are now recovered. However, Ubuntu lets us know visually that we can’t use these files yet. That’s because the ntfsundelete program saves the files as the “root” user, not the “ubuntu” user. We can verify this by typing the following in our terminal window: ls –l We want these three files to be owned by ubuntu, not root. To do this, enter the following in the terminal window: sudo chown ubuntu <Files> If the current folder has other files in it, you may not want to change their owner to ubuntu. However, in our case, we only have these three files in this folder, so we will use the * wildcard to change the owner of all three files. sudo chown ubuntu * The files now look normal, and we can do whatever we want with them. Hopefully you won’t need to use this tip, but if you do, ntfsundelete is a nice command-line utility. It doesn’t have a fancy GUI like many of the similar Windows programs, but it is a powerful tool that can recover your files quickly. See ntfsundelete’s manual page for more detailed usage information Similar Articles Productive Geek Tips Reset Your Ubuntu Password Easily from the Live CDUse Ubuntu Live CD to Backup Files from Your Dead Windows ComputerCreate a Bootable Ubuntu 9.10 USB Flash DriveCreate a Bootable Ubuntu USB Flash Drive the Easy WayGuide to Using Check Disk in Windows Vista TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 PCmover Professional Windows 7 Easter Theme YoWindoW, a real time weather screensaver Optimize your computer the Microsoft way Stormpulse provides slick, real time weather data Geek Parents – Did you try Parental Controls in Windows 7? Change DNS servers on the fly with DNS Jumper

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  • Persistent static routes fail on MacOS 10.6.5 startup!

    - by verbalicious
    I'm unable to get static routes to persist a reboot on Mac OS 10.6.5. I've tried all of the methods prescribed in Google search results, and previous posts on this site. I've tried manually creating a launchd daemon, and used RouteSplit's launchd daemon to no avail. It's clear that the interface is not ready when these methods attempt to apply the route. This workstation in question is getting its IP from DHCP and probably hasn't gotten its DHCP lease when the command runs. We're able to apply the route by hand when logged in, but not through startup methods. Is there another way to apply this route by sneaking the command into something later, but before the login window appears to the user? Here is some relevant log info from system.log. You can see the "route: writing to routing socket: Network is unreachable" errors where my launchd script fires off. I've tried adding extra "sleep" and "ipconfig waitall" statements later in the script but this doesn't fly. Dec 15 19:30:41 localhost com.apple.launchd[1]: *** launchd[1] has started up. *** Dec 15 19:30:45 localhost mDNSResponder[18]: mDNSResponder mDNSResponder-258.13 (Oct 8 2010 17:10:30) starting Dec 15 19:30:47 localhost configd[15]: bootp_session_transmit: bpf_write(en1) failed: Network is down (50) Dec 15 19:30:47 localhost configd[15]: DHCP en1: INIT transmit failed Dec 15 19:30:47 localhost configd[15]: network configuration changed. Dec 15 19:30:47 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:30:47 Administrators-MacBook-Pro blued[16]: Apple Bluetooth daemon started Dec 15 19:30:52 Administrators-MacBook-Pro syslog[67]: routes.sh: Starting RouteSplit Dec 15 19:30:53 Administrators-MacBook-Pro com.apple.usbmuxd[41]: usbmuxd-207 built for iTunesTenOne on Oct 19 2010 at 13:50:35, running 64 bit Dec 15 19:30:54 Administrators-MacBook-Pro /System/Library/CoreServices/loginwindow.app/Contents/MacOS/loginwindow[50]: Login Window Application Started Dec 15 19:30:55 Administrators-MacBook-Pro bootlog[61]: BOOT_TIME: 1292459441 0 Dec 15 19:30:55 Administrators-MacBook-Pro syslog[86]: routes.sh: static route 192.168.0.0/23 192.168.2.2 Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: route: writing to routing socket: Network is unreachable Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: add net 192.168.0.0: gateway 192.168.2.2: Network is unreachable Dec 15 19:30:57 Administrators-MacBook-Pro org.apache.httpd[38]: httpd: Could not reliably determine the server's fully qualified domain name, using Administrators-MacBook-Pro.local for ServerName Dec 15 19:30:58 Administrators-MacBook-Pro loginwindow[50]: Login Window Started Security Agent Dec 15 19:30:58 Administrators-MacBook-Pro WindowServer[89]: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:30:58 Administrators-MacBook-Pro com.apple.WindowServer[89]: Wed Dec 15 19:30:58 Administrators-MacBook-Pro.local WindowServer[89] <Error>: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:31:18 Administrators-MacBook-Pro configd[15]: network configuration changed. Dec 15 19:31:19 administrators-macbook-pro configd[15]: setting hostname to "administrators-macbook-pro.local" Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[121]: /usr/libexec/ntpd-wrapper: scutil key State:/Network/Global/DNS not present after 30 seconds Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp options: a=2 v=1 e=0.100 E=5.000 P=2147483647.000 Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: d=15 c=5 x=0 op=1 l=/var/run/sntp.pid f= time.apple.com Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp: getaddrinfo(hostname, ntp) failed with nodename nor servname provided, or not known Dec 15 19:31:27 administrators-macbook-pro configd[15]: network configuration changed. Dec 15 19:31:27 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:31:27 Administrators-MacBook-Pro ntpd[37]: Cannot find existing interface for address 17.151.16.20 Dec 15 19:31:27 Administrators-MacBook-Pro ntpd_initres[125]: ntpd indicates no data available! Dec 15 19:31:31 Administrators-MacBook-Pro sshd[128]: USER_PROCESS: 133 ttys000 Dec 15 19:31:37 Administrators-MacBook-Pro sudo[138]: administrator : TTY=ttys000 ; PWD=/Users/administrator ; USER=root ; COMMAND=/usr/bin/less /var/log/system.log ``You can see the following line in /var/log/kernel.log that shows the en0 interface coming up: Dec 15 19:30:51 Administrators-MacBook-Pro kernel[0]: Ethernet [AppleBCM5701Ethernet]: Link up on en0, 1-Gigabit, Full-duplex, No flow-control, Debug [796d,0f01,0de1,0300,c1e1,3800]

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  • LLBLGen Pro feature highlights: grouping model elements

    - by FransBouma
    (This post is part of a series of posts about features of the LLBLGen Pro system) When working with an entity model which has more than a few entities, it's often convenient to be able to group entities together if they belong to a semantic sub-model. For example, if your entity model has several entities which are about 'security', it would be practical to group them together under the 'security' moniker. This way, you could easily find them back, yet they can be left inside the complete entity model altogether so their relationships with entities outside the group are kept. In other situations your domain consists of semi-separate entity models which all target tables/views which are located in the same database. It then might be convenient to have a single project to manage the complete target database, yet have the entity models separate of each other and have them result in separate code bases. LLBLGen Pro can do both for you. This blog post will illustrate both situations. The feature is called group usage and is controllable through the project settings. This setting is supported on all supported O/R mapper frameworks. Situation one: grouping entities in a single model. This situation is common for entity models which are dense, so many relationships exist between all sub-models: you can't split them up easily into separate models (nor do you likely want to), however it's convenient to have them grouped together into groups inside the entity model at the project level. A typical example for this is the AdventureWorks example database for SQL Server. This database, which is a single catalog, has for each sub-group a schema, however most of these schemas are tightly connected with each other: adding all schemas together will give a model with entities which indirectly are related to all other entities. LLBLGen Pro's default setting for group usage is AsVisualGroupingMechanism which is what this situation is all about: we group the elements for visual purposes, it has no real meaning for the model nor the code generated. Let's reverse engineer AdventureWorks to an entity model. By default, LLBLGen Pro uses the target schema an element is in which is being reverse engineered, as the group it will be in. This is convenient if you already have categorized tables/views in schemas, like which is the case in AdventureWorks. Of course this can be switched off, or corrected on the fly. When reverse engineering, we'll walk through a wizard which will guide us with the selection of the elements which relational model data should be retrieved, which we can later on use to reverse engineer to an entity model. The first step after specifying which database server connect to is to select these elements. below we can see the AdventureWorks catalog as well as the different schemas it contains. We'll include all of them. After the wizard completes, we have all relational model data nicely in our catalog data, with schemas. So let's reverse engineer entities from the tables in these schemas. We select in the catalog explorer the schemas 'HumanResources', 'Person', 'Production', 'Purchasing' and 'Sales', then right-click one of them and from the context menu, we select Reverse engineer Tables to Entity Definitions.... This will bring up the dialog below. We check all checkboxes in one go by checking the checkbox at the top to mark them all to be added to the project. As you can see LLBLGen Pro has already filled in the group name based on the schema name, as this is the default and we didn't change the setting. If you want, you can select multiple rows at once and set the group name to something else using the controls on the dialog. We're fine with the group names chosen so we'll simply click Add to Project. This gives the following result:   (I collapsed the other groups to keep the picture small ;)). As you can see, the entities are now grouped. Just to see how dense this model is, I've expanded the relationships of Employee: As you can see, it has relationships with entities from three other groups than HumanResources. It's not doable to cut up this project into sub-models without duplicating the Employee entity in all those groups, so this model is better suited to be used as a single model resulting in a single code base, however it benefits greatly from having its entities grouped into separate groups at the project level, to make work done on the model easier. Now let's look at another situation, namely where we work with a single database while we want to have multiple models and for each model a separate code base. Situation two: grouping entities in separate models within the same project. To get rid of the entities to see the second situation in action, simply undo the reverse engineering action in the project. We still have the AdventureWorks relational model data in the catalog. To switch LLBLGen Pro to see each group in the project as a separate project, open the Project Settings, navigate to General and set Group usage to AsSeparateProjects. In the catalog explorer, select Person and Production, right-click them and select again Reverse engineer Tables to Entities.... Again check the checkbox at the top to mark all entities to be added and click Add to Project. We get two groups, as expected, however this time the groups are seen as separate projects. This means that the validation logic inside LLBLGen Pro will see it as an error if there's e.g. a relationship or an inheritance edge linking two groups together, as that would lead to a cyclic reference in the code bases. To see this variant of the grouping feature, seeing the groups as separate projects, in action, we'll generate code from the project with the two groups we just created: select from the main menu: Project -> Generate Source-code... (or press F7 ;)). In the dialog popping up, select the target .NET framework you want to use, the template preset, fill in a destination folder and click Start Generator (normal). This will start the code generator process. As expected the code generator has simply generated two code bases, one for Person and one for Production: The group name is used inside the namespace for the different elements. This allows you to add both code bases to a single solution and use them together in a different project without problems. Below is a snippet from the code file of a generated entity class. //... using System.Xml.Serialization; using AdventureWorks.Person; using AdventureWorks.Person.HelperClasses; using AdventureWorks.Person.FactoryClasses; using AdventureWorks.Person.RelationClasses; using SD.LLBLGen.Pro.ORMSupportClasses; namespace AdventureWorks.Person.EntityClasses { //... /// <summary>Entity class which represents the entity 'Address'.<br/><br/></summary> [Serializable] public partial class AddressEntity : CommonEntityBase //... The advantage of this is that you can have two code bases and work with them separately, yet have a single target database and maintain everything in a single location. If you decide to move to a single code base, you can do so with a change of one setting. It's also useful if you want to keep the groups as separate models (and code bases) yet want to add relationships to elements from another group using a copy of the entity: you can simply reverse engineer the target table to a new entity into a different group, effectively making a copy of the entity. As there's a single target database, changes made to that database are reflected in both models which makes maintenance easier than when you'd have a separate project for each group, with its own relational model data. Conclusion LLBLGen Pro offers a flexible way to work with entities in sub-models and control how the sub-models end up in the generated code.

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  • Persistent static routes fail on MacOS 10.6.5 startup!

    - by verbalicious
    I'm unable to get static routes to persist a reboot on Mac OS 10.6.5. I've tried all of the methods prescribed in Google search results, and previous posts on this site. I've tried manually creating a launchd daemon, and used RouteSplit's launchd daemon to no avail. It's clear that the interface is not ready when these methods attempt to apply the route. This workstation in question is getting its IP from DHCP and probably hasn't gotten its DHCP lease when the command runs. We're able to apply the route by hand when logged in, but not through startup methods. Is there another way to apply this route by sneaking the command into something later, but before the login window appears to the user? Here is some relevant log info from system.log. You can see the "route: writing to routing socket: Network is unreachable" errors where my launchd script fires off. I've tried adding extra "sleep" and "ipconfig waitall" statements later in the script but this doesn't fly. Dec 15 19:30:41 localhost com.apple.launchd[1]: *** launchd[1] has started up. *** Dec 15 19:30:45 localhost mDNSResponder[18]: mDNSResponder mDNSResponder-258.13 (Oct 8 2010 17:10:30) starting Dec 15 19:30:47 localhost configd[15]: bootp_session_transmit: bpf_write(en1) failed: Network is down (50) Dec 15 19:30:47 localhost configd[15]: DHCP en1: INIT transmit failed Dec 15 19:30:47 localhost configd[15]: network configuration changed. Dec 15 19:30:47 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:30:47 Administrators-MacBook-Pro blued[16]: Apple Bluetooth daemon started Dec 15 19:30:52 Administrators-MacBook-Pro syslog[67]: routes.sh: Starting RouteSplit Dec 15 19:30:53 Administrators-MacBook-Pro com.apple.usbmuxd[41]: usbmuxd-207 built for iTunesTenOne on Oct 19 2010 at 13:50:35, running 64 bit Dec 15 19:30:54 Administrators-MacBook-Pro /System/Library/CoreServices/loginwindow.app/Contents/MacOS/loginwindow[50]: Login Window Application Started Dec 15 19:30:55 Administrators-MacBook-Pro bootlog[61]: BOOT_TIME: 1292459441 0 Dec 15 19:30:55 Administrators-MacBook-Pro syslog[86]: routes.sh: static route 192.168.0.0/23 192.168.2.2 Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: route: writing to routing socket: Network is unreachable Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: add net 192.168.0.0: gateway 192.168.2.2: Network is unreachable Dec 15 19:30:57 Administrators-MacBook-Pro org.apache.httpd[38]: httpd: Could not reliably determine the server's fully qualified domain name, using Administrators-MacBook-Pro.local for ServerName Dec 15 19:30:58 Administrators-MacBook-Pro loginwindow[50]: Login Window Started Security Agent Dec 15 19:30:58 Administrators-MacBook-Pro WindowServer[89]: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:30:58 Administrators-MacBook-Pro com.apple.WindowServer[89]: Wed Dec 15 19:30:58 Administrators-MacBook-Pro.local WindowServer[89] <Error>: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:31:18 Administrators-MacBook-Pro configd[15]: network configuration changed. Dec 15 19:31:19 administrators-macbook-pro configd[15]: setting hostname to "administrators-macbook-pro.local" Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[121]: /usr/libexec/ntpd-wrapper: scutil key State:/Network/Global/DNS not present after 30 seconds Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp options: a=2 v=1 e=0.100 E=5.000 P=2147483647.000 Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: d=15 c=5 x=0 op=1 l=/var/run/sntp.pid f= time.apple.com Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp: getaddrinfo(hostname, ntp) failed with nodename nor servname provided, or not known Dec 15 19:31:27 administrators-macbook-pro configd[15]: network configuration changed. Dec 15 19:31:27 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:31:27 Administrators-MacBook-Pro ntpd[37]: Cannot find existing interface for address 17.151.16.20 Dec 15 19:31:27 Administrators-MacBook-Pro ntpd_initres[125]: ntpd indicates no data available! Dec 15 19:31:31 Administrators-MacBook-Pro sshd[128]: USER_PROCESS: 133 ttys000 Dec 15 19:31:37 Administrators-MacBook-Pro sudo[138]: administrator : TTY=ttys000 ; PWD=/Users/administrator ; USER=root ; COMMAND=/usr/bin/less /var/log/system.log ``You can see the following line in /var/log/kernel.log that shows the en0 interface coming up: Dec 15 19:30:51 Administrators-MacBook-Pro kernel[0]: Ethernet [AppleBCM5701Ethernet]: Link up on en0, 1-Gigabit, Full-duplex, No flow-control, Debug [796d,0f01,0de1,0300,c1e1,3800]

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  • nginx: problem configuring a proxy_pass

    - by Ofer Bar
    I'm converting a web app from apache to nginx. In apache's httpd.conf I have: ProxyPass /proxy/ http:// ProxyPassReverse /proxy/ http:// The idea is the client send this url: http://web-server-domain/proxy/login-server-addr/loginUrl.php?user=xxx&pass=yyy and the web server calls: http://login-server-addr/loginUrl.php?user=xxx&pass=yyy My nginx.conf is attached below and it is not working. At the moment it looks like it is calling the server, but returning an application error. This seems promising but any attempt to debug this failed! I can't trace any of the calls as nginx refuses to place them in the error file. Also, placing echo statement on the login server did not help either which is weird. The nginx documentation isn't very helpful about this. Any suggestion on how to configure a proxy_pass? Thanks! user nginx; worker_processes 1; #error_log /var/log/nginx/error.log; error_log /var/log/nginx/error.log notice; #error_log /var/log/nginx/error.log info; pid /var/run/nginx.pid; events { worker_connections 1024; } http { include /etc/nginx/mime.types; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log /var/log/nginx/access.log main; sendfile on; #tcp_nopush on; #keepalive_timeout 0; keepalive_timeout 65; #gzip on; # # The default server # server { rewrite_log on; listen 80; server_name _; #charset koi8-r; #access_log logs/host.access.log main; #root /var/www/live/html; index index.php index.html index.htm; location ~ ^/proxy/(.*$) { #location /proxy/ { # rewrite ^/proxy(.*) http://$1 break; proxy_set_header X-Forwarded-Host $host; proxy_set_header X-Forwarded-Server $host; proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_buffering off; proxy_pass http://$1; #proxy_pass "http://173.231.134.36/messages_2.7.0/loginUser.php?userID=ofer.fly%40gmail.com&password=y4HTD93vrshMNcy2Qr5ka7ia0xcaa389f4885f59c9"; break; } location / { root /var/www/live/html; #if ( $uri ~ ^/proxy/(.*) ) { # proxy_pass http://$1; # break; #} #try_files $uri $uri/ /index.php; } error_page 404 /404.html; #location = /404.html { # root /usr/share/nginx/html; #} # redirect server error pages to the static page /50x.html # #error_page 500 502 503 504 /50x.html; #location = /50x.html { # root /usr/share/nginx/html; #} # proxy the PHP scripts to Apache listening on 127.0.0.1:80 # #location ~ \.php$ { # proxy_pass http://127.0.0.1; #} # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # location ~ \.php$ { #root html; fastcgi_pass 127.0.0.1:9000; fastcgi_index index.php; #fastcgi_param SCRIPT_FILENAME /usr/share/nginx/html$fastcgi_script_name; fastcgi_param SCRIPT_FILENAME /var/www/live/html$fastcgi_script_name; include fastcgi_params; } # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # #location ~ /\.ht { # deny all; #} } # Load config files from the /etc/nginx/conf.d directory include /etc/nginx/conf.d/*.conf; }

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