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  • Practices for domain models in Javascript (with frameworks)

    - by AndyBursh
    This is a question I've to-and-fro'd with for a while, and searched for and found nothing on: what're the accepted practices surrounding duplicating domain models in Javascript for a web application, when using a framework like Backbone or Knockout? Given a web application of a non-trivial size with a set of domain models on the server side, should we duplicate these models in the web application (see the example at the bottom)? Or should we use the dynamic nature to load these models from the server? To my mind, the arguments for duplicating the models are in easing validation of fields, ensuring that fields that expected to be present are in fact present etc. My approach is to treat the client-side code like an almost separate application, doing trivial things itself and only relying on the server for data and complex operations (which require data the client-side doesn't have). I think treating the client-side code like this is akin to separation between entities from an ORM and the models used with the view in the UI layer: they may have the same fields and relate to the same domain concept, but they're distinct things. On the other hand, it seems to me that duplicating these models on the server side is a clear violation of DRY and likely to lead to differing results on the client- and server-side (where one piece gets updated but the other doesn't). To avoid this violation of DRY we can simply use Javascripts dynamism to get the field names and data from the server as and when they're neeed. So: are there any accepted guidelines around when (and when not) to repeat yourself in these situations? Or this a purely subjective thing, based on the project and developer(s)? Example Server-side model class M { int A DateTime B int C int D = (A*C) double SomeComplexCalculation = ServiceLayer.Call(); } Client-side model function M(){ this.A = ko.observable(); this.B = ko.observable(); this.C = ko.observable(); this.D = function() { return A() * C(); } this.SomeComplexCalculation = ko.observalbe(); return this; }l M.GetComplexValue = function(){ this.SomeComplexCalculation(Ajax.CallBackToServer()); }; I realise this question is quite similar to this one, but I think this is more about almost wholly untying the web application from the server, where that question is about doing this only in the case of complex calculation.

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  • Validation and Error Generation when using the Data Mapper Pattern

    - by AndyPerlitch
    I am working on saving state of an object to a database using the data mapper pattern, but I am looking for suggestions/guidance on the validation and error message generation step (step 4 below). Here are the general steps as I see them for doing this: (1) The data mapper is used to get current info (assoc array) about the object in db: +=====================================================+ | person_id | name | favorite_color | age | +=====================================================+ | 1 | Andy | Green | 24 | +-----------------------------------------------------+ mapper returns associative array, eg. Person_Mapper::getPersonById($id) : $person_row = array( 'person_id' => 1, 'name' => 'Andy', 'favorite_color' => 'Green', 'age' => '24', ); (2) the Person object constructor takes this array as an argument, populating its fields. class Person { protected $person_id; protected $name; protected $favorite_color; protected $age; function __construct(array $person_row) { $this->person_id = $person_row['person_id']; $this->name = $person_row['name']; $this->favorite_color = $person_row['favorite_color']; $this->age = $person_row['age']; } // getters and setters... public function toArray() { return array( 'person_id' => $this->person_id, 'name' => $this->name, 'favorite_color' => $this->favorite_color, 'age' => $this->age, ); } } (3a) (GET request) Inputs of an HTML form that is used to change info about the person is populated using Person::getters <form> <input type="text" name="name" value="<?=$person->getName()?>" /> <input type="text" name="favorite_color" value="<?=$person->getFavColor()?>" /> <input type="text" name="age" value="<?=$person->getAge()?>" /> </form> (3b) (POST request) Person object is altered with the POST data using Person::setters $person->setName($_POST['name']); $person->setFavColor($_POST['favorite_color']); $person->setAge($_POST['age']); *(4) Validation and error message generation on a per-field basis - Should this take place in the person object or the person mapper object? - Should data be validated BEFORE being placed into fields of the person object? (5) Data mapper saves the person object (updates row in the database): $person_mapper->savePerson($person); // the savePerson method uses $person->toArray() // to get data in a more digestible format for the // db gateway used by person_mapper Any guidance, suggestions, criticism, or name-calling would be greatly appreciated.

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  • The type of programmer I want to be [closed]

    - by Aventinus_
    I'm an undergraduate Software Engineer student, although I've decided that pure programming is what I want to do for the rest of my life. The thing is that programming is a vast field and although most of its aspects are extremely interesting, soon or later I'll have to choose one (?) to focus on. I have several ideas on small projects I'd like to develop this summer, having in mind that this will gain me some experience and, in the best scenario, some cash. But the most important reason I'd like to develop something close to “professional” is to give myself direction on what I want to do as a programmer. One path is that of the Web Programmer. I enjoy PHP and MySQL, as well as HTML and CSS, although I don't really like ASP.NET. I can see myself writing web apps, using the above technologies, as well as XML and Javascript. I also have a neat idea on a Facebook app. The other path is that of the Desktop Programmer. This is a little more complicated cause I really-really enjoy high level languages such as Java and Python but not the low level ones, such as C. I use both Linux and Windows for the last 6 years and I like their latest DEs (meaning Gnome Shell and Metro). I can see myself writing desktop applications for both OSs as long as it means high level programming. Ideally I'd like being able to help the development of GNOME. The last path that interests me is the path of the Smartphone Programmer. I have created some sample applications on Android and due to Java I found it a quite interesting experience. I can also see myself as an independent smartphone developer. These 3 paths seem equally interesting at the moment due to the shallowness of my experience, I guess. I know that I should spend time with all of them and then choose the right one for me but I'd like to know what are the pros and cons in terms of learning curve, fun, job finding and of course financial rewards with each of these paths. I have fair or basic understanding of the languages/technologies I described earlier and this question will help me choose where to focus, at least for now.

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  • LibGDX - Textures rendering at wrong position

    - by ACluelessGuy
    Update 2: Let me further explain my problem since I think that i didn't make it clear enough: The Y-coordinates on the bottom of my screen should be 0. Instead it is the height of my screen. That means the "higher" i touch/click the screen the less my y-coordinate gets. Above that the origin is not inside my screen, atleast not the 0 y-coordinate. Original post: I'm currently developing a tower defence game for fun by using LibGDX. There are places on my map where the player is or is not allowed to put towers on. So I created different ArrayLists holding rectangles representing a tile on my map. (towerPositions) for(int i = 0; i < map.getLayers().getCount(); i++) { curLay = (TiledMapTileLayer) map.getLayers().get(i); //For all Cells of current Layer for(int k = 0; k < curLay.getWidth(); k++) { for(int j = 0; j < curLay.getHeight(); j++) { curCell = curLay.getCell(k, j); //If there is a actual cell if(curCell != null) { tileWidth = curLay.getTileWidth(); tileHeight = curLay.getTileHeight(); xTileKoord = tileWidth*k; yTileKoord = tileHeight*j; switch(curLay.getName()) { //If layer named "TowersAllowed" picked case "TowersAllowed": towerPositions.add(new Rectangle(xTileKoord, yTileKoord, tileWidth, tileHeight)); // ... AND SO ON If the player clicks on a "allowed" field later on he has the opportunity to build a tower of his coice via a menu. Now here is the problem: The towers render, but they render at wrong position. (They appear really random on the map, no certain pattern for me) for(Rectangle curRect : towerPositions) { if(curRect.contains(xCoord, yCoord)) { //Using a certain tower in this example (left the menu out if(gameControl.createTower("towerXY")) { //RenderObject is just a class holding the Texture and x/y coordinates renderList.add(new RenderObject(new Texture(Gdx.files.internal("TowerXY.png")), curRect.x, curRect.y)); } } } Later on i render it: game.batch.begin(); for(int i = 0; i < renderList.size() ; i++) { game.batch.draw(renderList.get(i).myTexture, renderList.get(i).x, renderList.get(i).y); } game.batch.end(); regards

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  • Does OO, TDD, and Refactoring to Smaller Functions affect Speed of Code?

    - by Dennis
    In Computer Science field, I have noticed a notable shift in thinking when it comes to programming. The advice as it stands now is write smaller, more testable code refactor existing code into smaller and smaller chunks of code until most of your methods/functions are just a few lines long write functions that only do one thing (which makes them smaller again) This is a change compared to the "old" or "bad" code practices where you have methods spanning 2500 lines, and big classes doing everything. My question is this: when it call comes down to machine code, to 1s and 0s, to assembly instructions, should I be at all concerned that my class-separated code with variety of small-to-tiny functions generates too much extra overhead? While I am not exactly familiar with how OO code and function calls are handled in ASM in the end, I do have some idea. I assume that each extra function call, object call, or include call (in some languages), generate an extra set of instructions, thereby increasing code's volume and adding various overhead, without adding actual "useful" code. I also imagine that good optimizations can be done to ASM before it is actually ran on the hardware, but that optimization can only do so much too. Hence, my question -- how much overhead (in space and speed) does well-separated code (split up across hundreds of files, classes, and methods) actually introduce compared to having "one big method that contains everything", due to this overhead? UPDATE for clarity: I am assuming that adding more and more functions and more and more objects and classes in a code will result in more and more parameter passing between smaller code pieces. It was said somewhere (quote TBD) that up to 70% of all code is made up of ASM's MOV instruction - loading CPU registers with proper variables, not the actual computation being done. In my case, you load up CPU's time with PUSH/POP instructions to provide linkage and parameter passing between various pieces of code. The smaller you make your pieces of code, the more overhead "linkage" is required. I am concerned that this linkage adds to software bloat and slow-down and I am wondering if I should be concerned about this, and how much, if any at all, because current and future generations of programmers who are building software for the next century, will have to live with and consume software built using these practices. UPDATE: Multiple files I am writing new code now that is slowly replacing old code. In particular I've noted that one of the old classes was a ~3000 line file (as mentioned earlier). Now it is becoming a set of 15-20 files located across various directories, including test files and not including PHP framework I am using to bind some things together. More files are coming as well. When it comes to disk I/O, loading multiple files is slower than loading one large file. Of course not all files are loaded, they are loaded as needed, and disk caching and memory caching options exist, and yet still I believe that loading multiple files takes more processing than loading a single file into memory. I am adding that to my concern.

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  • Lazy Processing of Streams

    - by Giorgio
    I have the following problem scenario: I have a text file and I have to read it and split it into lines. Some lines might need to be dropped (according to criteria that are not fixed). The lines that are not dropped must be parsed into some predefined records. Records that are not valid must be dropped. Duplicate records may exist and, in such a case, they are consecutive. If duplicate / multiple records exist, only one item should be kept. The remaining records should be grouped according to the value contained in one field; all records belonging to the same group appear one after another (e.g. AAAABBBBCCDEEEFF and so on). The records of each group should be numbered (1, 2, 3, 4, ...). For each group the numbering starts from 1. The records must then be saved somewhere / consumed in the same order as they were produced. I have to implement this in Java or C++. My first idea was to define functions / methods like: One method to get all the lines from the file. One method to filter out the unwanted lines. One method to parse the filtered lines into valid records. One method to remove duplicate records. One method to group records and number them. The problem is that the data I am going to read can be too big and might not fit into main memory: so I cannot just construct all these lists and apply my functions one after the other. On the other hand, I think I do not need to fit all the data in main memory at once because once a record has been consumed all its underlying data (basically the lines of text between the previous record and the current record, and the record itself) can be disposed of. With the little knowledge I have of Haskell I have immediately thought about some kind of lazy evaluation, in which instead of applying functions to lists that have been completely computed, I have different streams of data that are built on top of each other and, at each moment, only the needed portion of each stream is materialized in main memory. But I have to implement this in Java or C++. So my question is which design pattern or other technique can allow me to implement this lazy processing of streams in one of these languages.

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  • Get Proactive With AutoVue Support!

    - by GrahamOracle
    v\:* {behavior:url(#default#VML);} o\:* {behavior:url(#default#VML);} w\:* {behavior:url(#default#VML);} .shape {behavior:url(#default#VML);} I’m pleased to announce the AutoVue “Get Proactive” page within the My Oracle Support portal. This new dynamic document links to valuable resources for AutoVue users, administrators, and integrators – Not only to remain up-to-date on Support and technical topics, but also to enhance your knowledge in planning for upgrade and maintenance activities for your AutoVue products. To access the AutoVue Get Proactive page, log into the MOS portal, search for note number 432.1 in the search field at the top-right, and once in the document select “Agile and AutoVue” from the dropdown (as shown in the following screenshot): The Get Proactive page is a working document, and we plan to include new resources as they become available. Therefore make sure to bookmark the document, and if you have any suggestions please post them using the ‘Add Comment’ feature within the document, or through the AutoVue Community. Normal 0 false false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0cm 5.4pt 0cm 5.4pt; mso-para-margin-top:0cm; mso-para-margin-right:0cm; mso-para-margin-bottom:10.0pt; mso-para-margin-left:0cm; line-height:115%; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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  • Is it smart to take a year off from school to get experience?

    - by user134147
    firstly I apologize if this question is not appropriate for the site, but I've seen other similar (though slightly deviant) questions on this sight before and I know the people here are the most qualified to answer my question. Anyways, I'm currently between my sophomore and junior years at a 4 year university, and after a bit of deliberation I've decided on computer science as a major (BA, by the way, as a BS would require me to stay at least an extra year the way our program is set up). I've been interested now in programming for a few months and I've developed a passion for it in a very short time. I began learning C++, migrating to Java recently when I learned my school focuses on this language. Now, I should mention that the concept of higher education has never sat well with me, so part of my motivation for wanting to take time off is to truly challenge myself and see what I can accomplish when I actually try at something. The autodidact in me finds it difficult to focus on my passions while trying to keep a high GPA in unrelated classes. However, I understand the times we live in and therefore would plan to complete my degree after this year. So my question is whether or not the skills I learn in a year off from college could justify the time off from school. Unfortunately, I don't believe I know enough yet to gain any professional experience (internship, etc.) so I would mostly focus my time on learning Java and another language, possibly Wordpress (to gain an understanding of web programming concepts as I have not yet decided what field I want to get into, and to make some money to fund my off-year), and to delve into security concepts, which also interest me. I'm hoping I could work on projects, such as simple applications or contributions to open source software during this time to enhance my resume once I do finish school, so I can find a job out of college easier. I do not want to be the new hire who knows nothing beyond the concepts of his Java textbooks. Does anyone have any input about these thoughts of mine, or any ideas for where I should focus my studies or how high I might set the bar for my work? Thanks a lot everyone!

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  • How to configure a longer version Number in artifactory

    - by claudine
    The version-numbers for our jars have to be longer them x.x.x. We would rather need x.x.x.x to integrate some old-fashioned self-made mechanism. This is, because we tag our software with x.x.x and as soon as we have a delivery to a customer one specific jar has to be build exactly at this point of time to fit to another backend, which communicates with our program. For that reason this one jar has the version 2.3.4.1, when generated and in next delivery of the same Version it is build and named 2.3.4.2. Now artifactory cannot handle this an doesn't save more than x.x.x.2 in some cases. So we thought of maybe edit the regular expression in the maven repository layout (see attached Screenshot) Because testing the path in the field below shows, that it cannot handle the version number. Of course for the rest of our jars still x.x.x has to work.. For Example here is the maven-metadata.xml <?xml version="1.0" encoding="UTF-8"?> <metadata> <groupId>com.firm</groupId> <artifactId>someid</artifactId> <version>1.5.1</version> <versioning> <latest>1.5.1</latest> <release>1.5.1</release> <versions> <version>1.4.62</version> </versions> <lastUpdated>20120926073942</lastUpdated> </versioning> </metadata> The folder structure looks like: someid 1.4.62 1.4.62.1 1.4.62.2 1.4.62.3 If we deploy an new artifact version (1.4.62.1), the maven-metadata.xml contains the 1.4.62.1 version. But the artifactory overrides the version number (1.4.62.x) to (1.4.62) after an unspecified time. It seems that the artifactory only support major, minor and revision numbers, and deletes the buildnumber. Now we looking for a solution do disable this behavior. We use the JFrog Artifactory version 2.5.0 (rev. 13086). Any ideas, maybe? Thanks in andvance

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  • Is the Observer pattern adequate for this kind of scenario?

    - by Omega
    I'm creating a simple game development framework with Ruby. There is a node system. A node is a game entity, and it has position. It can have children nodes (and one parent node). Children are always drawn relatively to their parent. Nodes have a @position field. Anyone can modify it. When such position is modified, the node must update its children accordingly to properly draw them relatively to it. @position contains a Point instance (a class with x and y properties, plus some other useful methods). I need to know when a node's @position's state changes, so I can tell the node to update its children. This is easy if the programmer does something like this: @node.position = Point.new(300,300) Because it is equivalent to calling this: # Code in the Node class def position=(newValue) @position = newValue update_my_children # <--- I know that the position changed end But, I'm lost when this happens: @node.position.x = 300 The only one that knows that the position changed is the Point instance stored in the @position property of the node. But I need the node to be notified! It was at this point that I considered the Observer pattern. Basically, Point is now observable. When a node's position property is given a new Point instance (through the assignment operator), it will stop observing the previous Point it had (if any), and start observing the new one. When a Point instance gets a state change, all observers (the node owning it) will be notified, so now my node can update its children when the position changes. A problem is when this happens: @someNode.position = @anotherNode.position This means that two nodes are observing the same point. If I change one of the node's position, the other would change as well. To fix this, when a position is assigned, I plan to create a new Point instance, copy the passed argument's x and y, and store my newly created point instead of storing the passed one. Another problem I fear is this: somePoint = @node.position somePoint.x = 500 This would, technically, modify @node's position. I'm not sure if anyone would be expecting that behavior. I'm under the impression that people see Point as some kind of primitive rather than an actual object. Is this approach even reasonable? Reasons I'm feeling skeptical: I've heard that the Observer pattern should be used with, well, many observers. Technically, in this scenario there is only one observer at a time. When assigning a node's position as another's (@someNode.position = @anotherNode.position), where I create a whole new instance rather than storing the passed point, it feels hackish, or even inefficient.

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  • Protobuf design patterns

    - by Monster Truck
    I am evaluating Google Protocol Buffers for a Java based service (but am expecting language agnostic patterns). I have two questions: The first is a broad general question: What patterns are we seeing people use? Said patterns being related to class organization (e.g., messages per .proto file, packaging, and distribution) and message definition (e.g., repeated fields vs. repeated encapsulated fields*) etc. There is very little information of this sort on the Google Protobuf Help pages and public blogs while there is a ton of information for established protocols such as XML. I also have specific questions over the following two different patterns: Represent messages in .proto files, package them as a separate jar, and ship it to target consumers of the service --which is basically the default approach I guess. Do the same but also include hand crafted wrappers (not sub-classes!) around each message that implement a contract supporting at least these two methods (T is the wrapper class, V is the message class (using generics but simplified syntax for brevity): public V toProtobufMessage() { V.Builder builder = V.newBuilder(); for (Item item : getItemList()) { builder.addItem(item); } return builder.setAmountPayable(getAmountPayable()). setShippingAddress(getShippingAddress()). build(); } public static T fromProtobufMessage(V message_) { return new T(message_.getShippingAddress(), message_.getItemList(), message_.getAmountPayable()); } One advantage I see with (2) is that I can hide away the complexities introduced by V.newBuilder().addField().build() and add some meaningful methods such as isOpenForTrade() or isAddressInFreeDeliveryZone() etc. in my wrappers. The second advantage I see with (2) is that my clients deal with immutable objects (something I can enforce in the wrapper class). One disadvantage I see with (2) is that I duplicate code and have to sync up my wrapper classes with .proto files. Does anyone have better techniques or further critiques on any of the two approaches? *By encapsulating a repeated field I mean messages such as this one: message ItemList { repeated item = 1; } message CustomerInvoice { required ShippingAddress address = 1; required ItemList = 2; required double amountPayable = 3; } instead of messages such as this one: message CustomerInvoice { required ShippingAddress address = 1; repeated Item item = 2; required double amountPayable = 3; } I like the latter but am happy to hear arguments against it.

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  • Basic Android game loop having issues

    - by WillDaBeast509
    I've set up a very basic game loop that should draw a circle, run 100 times, then draw another. I also have a text field that should display how many times the loop has ran. However, the screen seems to not update. It displays a different value for the tick count (different each time the app is ran) and simply stays there. After exiting the app, I get an error saying "Unfortunately, MyApp has stopped." Here is the relevant code: DrawView public class DrawView extends SurfaceView implements SurfaceHolder.Callback { Paint p = new Paint(); MainThread thread; private int y=0; public DrawView(Context c) { super(c); thread = new MainThread(this, getHolder()); thread.running = true; getHolder().addCallback(this); setFocusable(true); } public void draw(Canvas c) { if(c==null) return; //super.onDraw(c); c.drawColor(Color.WHITE); p.setColor(Color.RED); p.setTextSize(32); p.setTypeface(Typeface.SANS_SERIF); c.drawCircle(getWidth()/2-100,getHeight()/2, 50, p); c.drawText("y = " + y, 50, 50, p); if(y>=100) { Log.i("DRAW", "drawing circle"); c.drawCircle(getWidth()/2+100,getHeight()/2, 50, p); } else y++; Log.i("INFO", "y = " + y); } @Override public boolean onTouchEvent(MotionEvent event) { return true; } public void onDraw(Canvas c){} public void surfaceCreated(SurfaceHolder p1) { thread.start(); } public void surfaceChanged(SurfaceHolder p1, int p2, int p3, int p4) { // TODO: Implement this method } public void surfaceDestroyed(SurfaceHolder p1) { thread.running = false; boolean retry = true; while (retry) { try { thread.join(); retry = false; } catch (InterruptedException e) { Log.i("EX", "cathing exception"); } } } } MainThread public class MainThread extends Thread { private DrawView page; private SurfaceHolder holder; public boolean running; public MainThread(DrawView p, SurfaceHolder h) { super(); page = p; holder = h; } @Override public void run() { while(running) { Canvas c = holder.lockCanvas(); page.draw(c); holder.unlockCanvasAndPost(c); } } } Here is an example log outupt: http://pastebin.com/tM9dUPuk It counts the number of ticks correctly and should draw the second circle, but the screen looks like its not updating. After closing the app, the log continues to run and keep outputting "y = 100 drawing circle" until it crashes and shows the error report. What is going on and how can I fix these two problems?

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  • Are we queueing and serializing properly?

    - by insta
    We process messages through a variety of services (one message will touch probably 9 services before it's done, each doing a specific IO-related function). Right now we have a combination of the worst-case (XML data contract serialization) and best-case (in-memory MSMQ) for performance. The nature of the message means that our serialized data ends up about 12-15 kilobytes, and we process about 4 million messages per week. Persistent messages in MSMQ were too slow for us, and as the data grows we are feeling the pressure from MSMQ's memory-mapped files. The server is at 16GB of memory usage and growing, just for queueing. Performance also suffers when the memory usage is high, as the machine starts swapping. We're already doing the MSMQ self-cleanup behavior. I feel like there's a part we're doing wrong here. I tried using RavenDB to persist the messages and just queueing an identifier, but the performance there was very slow (1000 messages per minute, at best). I'm not sure if that's a result of using the development version or what, but we definitely need a higher throughput[1]. The concept worked very well in theory but performance was not up to the task. The usage pattern has one service acting as a router, which does all reads. The other services will attach information based on their 3rd party hook, and forward back to the router. Most objects are touched 9-12 times, although about 10% are forced to loop around in this system for awhile until the 3rd parties respond appropriately. The services right now account for this and have appropriate sleeping behaviors, as we utilize the priority field of the message for this reason. So, my question, is what is an ideal stack for message passing between discrete-but-LAN'ed machines in a C#/Windows environment? I would normally start with BinaryFormatter instead of XML serialization, but that's a rabbit hole if a better way is to offload serialization to a document store. Hence, my question. [1]: The nature of our business means the sooner we process messages, the more money we make. We've empirically proven that processing a message later in the week means we are less likely to make that money. While performance of "1000 per minute" sounds plenty fast, we really need that number upwards of 10k/minute. Just because I'm giving numbers in messages per week doesn't mean we have a whole week to process those messages.

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  • Architecture or Pattern for handling properties with custom setter/getter?

    - by Shelby115
    Current Situation: I'm doing a simple MVC site for keeping journals as a personal project. My concern is I'm trying to keep the interaction between the pages and the classes simplistic. Where I run into issues is the password field. My setter encrypts the password, so the getter retrieves the encrypted password. public class JournalBook { private IEncryptor _encryptor { get; set; } private String _password { get; set; } public Int32 id { get; set; } public String name { get; set; } public String description { get; set; } public String password { get { return this._password; } set { this.setPassword(this._password, value, value); } } public List<Journal> journals { get; set; } public DateTime created { get; set; } public DateTime lastModified { get; set; } public Boolean passwordProtected { get { return this.password != null && this.password != String.Empty; } } ... } I'm currently using model-binding to submit changes or create new JournalBooks (like below). The problem arises that in the code below book.password is always null, I'm pretty sure this is because of the custom setter. [HttpPost] public ActionResult Create(JournalBook book) { // Create the JournalBook if not null. if (book != null) this.JournalBooks.Add(book); return RedirectToAction("Index"); } Question(s): Should I be handling this not in the property's getter/setter? Is there a pattern or architecture that allows for model-binding or another simple method when properties need to have custom getters/setters to manipulate the data? To summarize, how can I handle the password storing with encryption such that I have the following, Robust architecture I don't store the password as plaintext. Submitting a new or modified JournalBook is as easy as default model-binding (or close to it).

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  • Series On Embedded Development (Part 2) - Build-Time Optionality

    - by user12612705
    In this entry on embedded development, I'm going to discuss build-time optionality (BTO). BTO is the ability to subset your software at build-time so you only use what is needed. BTO typically pertains more to software providers rather then developers of final products. For example, software providers ship source products, frameworks or platforms which are used by developers to build other products. If you provide a source product, you probably don't have to do anything to support BTO as the developers using your source will only use the source they need to build their product. If you provide a framework, then there are some things you can do to support BTO. Say you provide a Java framework which supports audio and video. If you provide this framework in a single JAR, then developers who only want audio are forced to ship their product with the video portion of your framework even though they aren't using it. In this case, support providing the framework in separate JARs...break the framework into an audio JAR and a video JAR and let the users of your framework decide which JARs to include in their product. Sometimes this is as simple as packaging, but if, for example, the video functionality is dependent on the audio functionality, it may require coding work to cleanly separate the two. BTO can also work at install-time, and this is sometimes overlooked. Let's say your building a phone application which can use Near Field Communications (NFC) if it's available on the phone, but it doesn't require NFC to work. Typically you'd write one app for all phones (saving you time)...both those that have NFC and those that don't, and just use NFC if it's there. However, for better efficiency, you can detect at install-time if the phone supports NFC and not install the NFC portion of your app if the phone doesn't support NFC. This requires that you write the app so it can run without the optional NFC code and that you write your install app so it can detect NFC and do the right thing at install-time. Supporting install-time optionality will save persistent footprint on the phone, something your customers will appreciate, your app "neighbors" will appreciate, and that you'll appreciate when they save static footprint for you. In the next article, I'll talk about runtime optionality.

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  • LWJGL: Camera distance from image plane?

    - by Rogem
    Let me paste some code before I ask the question... public static void createWindow(int[] args) { try { Display.setFullscreen(false); DisplayMode d[] = Display.getAvailableDisplayModes(); for (int i = 0; i < d.length; i++) { if (d[i].getWidth() == args[0] && d[i].getHeight() == args[1] && d[i].getBitsPerPixel() == 32) { displayMode = d[i]; break; } } Display.setDisplayMode(displayMode); Display.create(); } catch (Exception e) { e.printStackTrace(); System.exit(0); } } public static void initGL() { GL11.glEnable(GL11.GL_TEXTURE_2D); GL11.glShadeModel(GL11.GL_SMOOTH); GL11.glClearColor(0.0f, 0.0f, 0.0f, 0.0f); GL11.glClearDepth(1.0); GL11.glEnable(GL11.GL_DEPTH_TEST); GL11.glDepthFunc(GL11.GL_LEQUAL); GL11.glMatrixMode(GL11.GL_PROJECTION); GL11.glLoadIdentity(); GLU.gluPerspective(45.0f, (float) displayMode.getWidth() / (float) displayMode.getHeight(), 0.1f, 100.0f); GL11.glMatrixMode(GL11.GL_MODELVIEW); GL11.glHint(GL11.GL_PERSPECTIVE_CORRECTION_HINT, GL11.GL_NICEST); } So, with the camera and screen setup out of the way, I can now ask the actual question: How do I know what the camera distance is from the image plane? I also would like to know what the angle between the image plane's center normal and a line drawn from the middle of one of the edges to the camera position is. This will be used to consequently draw a vector from the camera's position through the player's click-coordinates to determine the world coordinates they clicked (or could've clicked). Also, when I set the camera coordinates, do I set the coordinates of the camera or do I set the coordinates of the image plane? Thank you for your help. EDIT: So, I managed to solve how to calculate the distance of the camera... Here's the relevant code... private static float getScreenFOV(int dim) { if (dim == 0) { float dist = (float) Math.tan((Math.PI / 2 - Math.toRadians(FOV_Y))/2) * 0.5f; float FOV_X = 2 * (float) Math.atan(getScreenRatio() * 0.5f / dist); return FOV_X; } else if (dim == 1) { return FOV_Y; } return 0; } FOV_Y is the Field of View that one defines in gluPerspective (float fovy in javadoc). This seems to be (and would logically be) for the height of the screen. Now I just need to figure out how to calculate that vector.

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  • What exactly is a X-YMailISG header?

    - by iainH
    Finally ... our emails are being seen by Yahoo! not as junk anymore. Hurray! However I notice that the Yahoo! receiving MTA adds in a X-YMailISG header. It's very large ... 2**10 bits? Now that I've invested too large a chunk of my waking life in crafting our email headers I'm curious to know what an X-YMailISG header is. Can anybody tell me? Does it pose any security / authenticity issues? There's very little intelligible from Google results. Background: After many days tweaking TXT records in our domain's DNS zone file for SPF and DKIM, I have at last succeeded in generating email from our Drupal site that Yahoo! no longer marks as X-YahooFilteredBulk and the excellent service [email protected] returns results that show the emails passing SPF, DKIM and Sender-ID checks and appearing to SpamAssassin as ham. Yahoo! even adds a Received-SPF: pass header. Useful links: http://www.goldfisch.at/knowwiki/howtos/dkim-filter http://old.openspf.org/wizard.html Strangely enough the SPF TXT record needed / allowed a blank key / name field in our registrar's DNS management panel whereas the DKIM record needed the {selector}._domainkey as the key /name of the DKIM strings.

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  • yum not working on EC2 Red Hat instance: Cannot retrieve repository metadata

    - by adev3
    For some reason yum has stopped working in my Amazon EC2 instance, located in the EU West sector. There seems to be something wrong with the path of the repo metadata, is this correct? I would be very grateful for any help, as my experience in this field is somewhat limited. Thank you very much. cat /etc/redhat-release: Red Hat Enterprise Linux Server release 6.2 (Santiago) yum repolist: Loaded plugins: amazon-id, rhui-lb, security https://rhui2-cds01.eu-west-1.aws.ce.redhat.com/pulp/repos//rhui-client-config/rhel/server/6/x86_64/os/repodata/repomd.xml: [Errno 14] PYCURL ERROR 22 - "The requested URL returned error: 401" Trying other mirror. https://rhui2-cds02.eu-west-1.aws.ce.redhat.com/pulp/repos//rhui-client-config/rhel/server/6/x86_64/os/repodata/repomd.xml: [Errno 14] PYCURL ERROR 22 - "The requested URL returned error: 401" Trying other mirror. repo id repo name status rhui-eu-west-1-client-config-server-6 Red Hat Update Infrastructure 2.0 Client Configuration Server 6 0 rhui-eu-west-1-rhel-server-releases Red Hat Enterprise Linux Server 6 (RPMs) 0 rhui-eu-west-1-rhel-server-releases-optional Red Hat Enterprise Linux Server 6 Optional (RPMs) 0 repolist: 0 yum update: (I needed to remove the base URLs below because of ServerFault's restrictions for new users) Loaded plugins: amazon-id, rhui-lb, security [same as base url 1 above]/pulp/repos//rhui-client-config/rhel/server/6/x86_64/os/repodata/repomd.xml: [Errno 14] PYCURL ERROR 22 - "The requested URL returned error: 401" Trying other mirror. [same as base url 2 above]/pulp/repos//rhui-client-config/rhel/server/6/x86_64/os/repodata/repomd.xml: [Errno 14] PYCURL ERROR 22 - "The requested URL returned error: 401" Trying other mirror. Error: Cannot retrieve repository metadata (repomd.xml) for repository: rhui-eu-west-1-client-config-server-6. Please verify its path and try again

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  • Zabbix Proxy not collecting data

    - by Jordan Eunson
    I have a working Zabbix 1.8.2 server collecting data for our office and our colo facility. However the link between the colo and office is flaky. What I'm trying to do is setup a proxy on the colo side to have a 1 hour cache and relay the data to our primary server at the office. Our zabbix server is compiled from source and uses a mysql database I've followed the instructions found in the zabbix documentation to compile the proxy using a sqlite3 database. I add the proxy to zabbix under Administration-DM-Proxies. The zabbix server "sees" the proxy because the "last seen" field is always under 60s. However when I assign a colo host to the proxy I stop receiving data from it. The colo host's zabbix_agentd.log file says this: 29343:20100622:124847 Timeout while answering request 29343:20100622:124847 Getting list of active checks failed. Will retry after 60 seconds The zabbix_proxy.log says this. 2041:20100622:123131.760 Deleted 0 records from history [0.000994 seconds] 2028:20100622:124131.671 Error while receiving answer from server [ZBX_TCP_READ() failed I also am unable to receive any SNMP data which is more important to me than the zabbix agent data. Has anyone had this problem before? Zabbix Server OS: CentOS5.4 Zabbix Server Build: 1.8.2 from source Zabbix Proxy OS: CentOS5.4 Zabbix Proxy Build: 1.8.2 from source P.S. The SQLite database on the zabbix proxy never gets any data written to it, it is identical to when I created it from the blank schema in zabbix-1.8.2/create/schema. (Yes I've checked the permissions)

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  • List of Hidden / Virtual Windows User Accounts

    - by Synetech inc.
    I’m trying to find a way to get a comprehensive list of user accounts on a Windows 7 system, including hidden ones. The User Accounts dialog (>control userpasswords2) only shows the normal user accounts, and even the Local User and Groups editor only shows normal user accounts and standard hidden/disabled ones like Administrator and Guest. The Select Users or Groups dialog has a Find Now button which which combines users and groups, but alas, it has the same contents as the LUG. I’m looking for a more comprehensive list that includes “super-hidden” / virtual user accounts like TrustedInstaller (or to be more accurate, NT Service\TrustedInstaller—notice the different “domain”). I checked HKLM\SOFTWARE\Microsoft\Windows NT\CurrentVersion\Winlogon\SpecialAccounts\UserList, but the SpecialAccounts key does not exist. I also checked HKLM\SOFTWARE\Microsoft\Windows NT\CurrentVersion\ProfileList, and while it does have the SystemProfile, LocalService, and NetworkService accounts listed, it does not have others (like TrustedInstaller and its ilk). TrustedInstaller specifically is a little confusing because it is a user, a service, and an executable file. I am using it as an example because it is “super hidden” in that it does not seem to be listed in any sort of user list. (As an experiment, I tried searching the whole registry for “trustedinstaller” to see if I could find a place where it is listed as a user, but found none.) To be clear, what I am looking for is a list of all accounts that can be used in a user input-field such as in permissions dialogs or as a runas argument.

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  • ubuntu 9.10 cups server cupsd.conf

    - by aaron
    i have a cups server running on ubuntu 9.10 on my home network. right now i can access it at 192.168.1.101:631, but when i try to access it at myservername.local:631 i get a 400 Bad Request. here's the relevant section from my current cupsd.conf: ServerName 192.168.1.101 # Only listen for connections from the local machine. Listen localhost:631 Listen /var/run/cups/cups.sock # any of the below 'Listen' directives all yield the same result Listen 192.168.1.101:631 #Listen *:631 #Listen myservername.local:631 # Show shared printers on the local network. Browsing On BrowseOrder allow,deny BrowseAllow all BrowseLocalProtocols CUPS dnssd BrowseAddress 192.168.1.255 # Default authentication type, when authentication is required... DefaultAuthType Basic # Restrict access to the server... <Location /> Order deny,allow Deny from All Allow from 127.0.0.1 Allow from 192.168.1.* </Location> # Restrict access to the admin pages... <Location /admin> Order deny,allow Deny from All #Allow from 127.0.0.1 #Allow from 192.168.1.* </Location> # Restrict access to configuration files... <Location /admin/conf> AuthType Default Require user @SYSTEM Order deny,allow Deny from All #Allow from 127.0.0.1 #Allow from 192.168.1.* </Location> i get the following in /var/log/cups/error_log: E [03/Jan/2010:18:33:41 -0600] Request from "192.168.1.100" using invalid Host: field "myservername.local:631" what do i need to do to be able to access the cups server at both 192.168.1.101:631 and myservername.local:631?

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  • Alter charset and collation in all columns in all tables in MySQL

    - by The Disintegrator
    I need to execute these statements in all tables for all columns. alter table table_name charset=utf8; alter table table_name alter column column_name charset=utf8; Is it possible to automate this in any way inside MySQL? I would prefer to avoid mysqldump Update: Richard Bronosky showed me the way :-) The query I needed to execute in every table: alter table DBname.DBfield CONVERT TO CHARACTER SET utf8 COLLATE utf8_general_ci; Crazy query to generate all other queries: SELECT distinct CONCAT( 'alter table ', TABLE_SCHEMA, '.', TABLE_NAME, ' CONVERT TO CHARACTER SET utf8 COLLATE utf8_general_ci;' ) FROM information_schema.COLUMNS WHERE TABLE_SCHEMA = 'DBname'; I only wanted to execute it in one database. It was taking too long to execute all in one pass. It turned out that it was generating one query per field per table. And only one query per table was necessary (distinct to the rescue). Getting the output on a file was how I realized it. How to generate the output to a file: mysql -B -N --user=user --password=secret -e "SELECT distinct CONCAT( 'alter table ', TABLE_SCHEMA, '.', TABLE_NAME, ' CONVERT TO CHARACTER SET utf8 COLLATE utf8_general_ci;' ) FROM information_schema.COLUMNS WHERE TABLE_SCHEMA = 'DBname';" > alter.sql And finally to execute all the queries: mysql --user=user --password=secret < alter.sql Thanks Richard. You're the man!

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  • Can't compile CentOS 5, Ruby 1.9.2 and OpenSSL 1.0.0c

    - by pstinnett
    I'm trying to install Ruby 1.9.2 on CentOS 5.5. I get through most of the make process, but when it tries to compile OpenSSL I get an error. Below is the errror outputted: compiling openssl make[1]: Entering directory `/sources/ruby-1.9.2-p136/ext/openssl' gcc -I. -I../../.ext/include/x86_64-linux -I../.././include -I../.././ext/openssl -DRUBY_EXTCONF_H=\"extconf.h\" -fPIC -O3 -ggdb -Wextra -Wno-unused-parameter -Wno-parentheses -Wpointer-arith -Wwrite-strings -Wno-missing-field-initializers -Wno-long-long -o ossl_x509.o -c ossl_x509.c In file included from ossl.h:201, from ossl_x509.c:11: openssl_missing.h:71: error: conflicting types for ‘HMAC_CTX_copy’ /usr/include/openssl/hmac.h:102: error: previous declaration of ‘HMAC_CTX_copy’ was here openssl_missing.h:95: error: conflicting types for ‘EVP_CIPHER_CTX_copy’ /usr/include/openssl/evp.h:459: error: previous declaration of ‘EVP_CIPHER_CTX_copy’ was here make[1]: *** [ossl_x509.o] Error 1 make[1]: Leaving directory `/sources/ruby-1.9.2-p136/ext/openssl' make: *** [mkmain.sh] Error 1 Any help would be greatly appreciated! I'm not a master at Linux by any means, but I was able to successfully install this version of Ruby on our dev server. Our live server is running a newer version of OpenSSL which I'm assuming is why it's breaking. Just not sure what the fix is!

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  • Any Recommendations for a Web Based Large File Transfer System?

    - by Glen Richards
    I'm looking for a server software product that: Allows my users to share large files with: The general public securely to 1 or more people (notification via email, optionally with a token that gives them x period of time to download) Allows anyone in the general public to share files with my users. Perhaps by invitation. Has to be user friendly enough to allow my users to use this with out having to bug me as the admin. It needs to be a system that we can install on our own server (we don't want shared data sitting on anyone else's server) A web based solution. Using some kind or secure comms channel would be good too, eg, ssh Files to share could be over 1 GB. I found the question below. WebDav does not sound user friendly enough: http://serverfault.com/questions/86878/recommendations-for-a-secure-and-simple-dropbox-system I've done a lot of searching, but I can't get the search terms right. There are too many services that provide this, but I want something we can install on our own server. A last resort would be to roll my own. Any ideas appreciated. Glen EDIT Sorry Tom and Jeff but Glen specifically says that he's looking for a 'product' so given that I specialise in this field thought that my expertise in this area may have been of use to him. I don't see how him writing services is going to be easy for him to maintain going forward (large IT admin overhead) or simple for his users and the general public to work with.

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  • How to configure postfix for per-sender SASL authentication

    - by Marwan
    I have two gmail accounts, and I want to configure my local postfix server as a client which does SASL authentication with smtp.gmail.com:587 with credentials that depend on the sender address. So, let's say that my gmail accounts are: [email protected] and [email protected]. If I sent a mail with [email protected] in the FROM header field, then postfix should use the credentials: [email protected]:psswd1 to do SASL authentication with gmail SMTP server. Similarly with [email protected], it should use [email protected]:passwd2. Sounds fairly simple. Well, I followed the postfix official documentation at http://www.postfix.org/SASL_README.html, and I ended up with the following relevant configurations: /etc/postfix/main.cf smtp_sasl_auth_enable = yes smtp_sasl_security_options = noanonymous smtp_sasl_password_maps = hash:/etc/postfix/sasl_passwd smtp_sender_dependent_authentication = yes sender_dependent_relayhost_maps = hash:/etc/postfix/sender_relay smtp_tls_security_level = secure smtp_tls_CAfile = /etc/ssl/certs/Equifax_Secure_CA.pem smtp_tls_CApath = /etc/ssl/certs smtp_tls_session_cache_database = btree:/etc/postfix/smtp_scache smtp_tls_session_cache_timeout = 3600s smtp_tls_loglevel = 1 tls_random_source = dev:/dev/urandom relayhost = smtp.gmail.com:587 /etc/postfix/sasl_passwd [email protected] [email protected]:passwd1 [email protected] [email protected]:passwd2 smtp.gmail.com:587 [email protected]:passwd1 /etc/postfix/sender_relay [email protected] smtp.gmail.com:587 [email protected] smtp.gmail.com:587 After I'm done with the configurations I did: $ postmap /etc/postfix/sasl_passwd $ postmap /etc/postfix/sender_relay $ /etc/init.d/postfix restart The problem is that when I send a mail from [email protected], the message ends up in the destination with sender address [email protected] and NOT [email protected], which means that postfix always ignores the per-sender configurations and send the mail using the default credentials (the third line in /etc/postfix/sasl_passwd above). I checked the configurations multiple times and even compared them to those in various blog posts addressing the same issue but found them to be more or less the same as mine. So, can anyone point me in the right direction, in case I'm missing something? Many thanks.

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