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  • Python: How best to parse a simple grammar?

    - by Rosarch
    Ok, so I've asked a bunch of smaller questions about this project, but I still don't have much confidence in the designs I'm coming up with, so I'm going to ask a question on a broader scale. I am parsing pre-requisite descriptions for a course catalog. The descriptions almost always follow a certain form, which makes me think I can parse most of them. From the text, I would like to generate a graph of course pre-requisite relationships. (That part will be easy, after I have parsed the data.) Some sample inputs and outputs: "CS 2110" => ("CS", 2110) # 0 "CS 2110 and INFO 3300" => [("CS", 2110), ("INFO", 3300)] # 1 "CS 2110, INFO 3300" => [("CS", 2110), ("INFO", 3300)] # 1 "CS 2110, 3300, 3140" => [("CS", 2110), ("CS", 3300), ("CS", 3140)] # 1 "CS 2110 or INFO 3300" => [[("CS", 2110)], [("INFO", 3300)]] # 2 "MATH 2210, 2230, 2310, or 2940" => [[("MATH", 2210), ("MATH", 2230), ("MATH", 2310)], [("MATH", 2940)]] # 3 If the entire description is just a course, it is output directly. If the courses are conjoined ("and"), they are all output in the same list If the course are disjoined ("or"), they are in separate lists Here, we have both "and" and "or". One caveat that makes it easier: it appears that the nesting of "and"/"or" phrases is never greater than as shown in example 3. What is the best way to do this? I started with PLY, but I couldn't figure out how to resolve the reduce/reduce conflicts. The advantage of PLY is that it's easy to manipulate what each parse rule generates: def p_course(p): 'course : DEPT_CODE COURSE_NUMBER' p[0] = (p[1], int(p[2])) With PyParse, it's less clear how to modify the output of parseString(). I was considering building upon @Alex Martelli's idea of keeping state in an object and building up the output from that, but I'm not sure exactly how that is best done. def addCourse(self, str, location, tokens): self.result.append((tokens[0][0], tokens[0][1])) def makeCourseList(self, str, location, tokens): dept = tokens[0][0] new_tokens = [(dept, tokens[0][1])] new_tokens.extend((dept, tok) for tok in tokens[1:]) self.result.append(new_tokens) For instance, to handle "or" cases: def __init__(self): self.result = [] # ... self.statement = (course_data + Optional(OR_CONJ + course_data)).setParseAction(self.disjunctionCourses) def disjunctionCourses(self, str, location, tokens): if len(tokens) == 1: return tokens print "disjunction tokens: %s" % tokens How does disjunctionCourses() know which smaller phrases to disjoin? All it gets is tokens, but what's been parsed so far is stored in result, so how can the function tell which data in result corresponds to which elements of token? I guess I could search through the tokens, then find an element of result with the same data, but that feel convoluted... What's a better way to approach this problem?

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  • How to dynamically change the content of a facet in a custom component?

    - by romaintaz
    Hello, Let's consider that I want to extend an existing JSF component, such as the <rich:datatable/>. My main requirement is to dynamically modify the content of a <f:facet>, to change its content. What is the best way to achieve that? Or where is the best place in the code to achieve that? In my faces-config.xml, I have the following declaration: <faces-config> ... <component> <component-type>my.component.dataTable</component-type> <component-class>my.project.component.table.MyHtmlDataTable</component-class> </component> ... <render-kit> <render-kit-id>HTML_BASIC</render-kit-id> <renderer> <component-family>org.richfaces.DataTable</component-family> <renderer-type>my.renderkit.dataTable</renderer-type> <renderer-class>my.project.component.table.MyDataTableRenderer</renderer-class> </renderer> ... Also, my my-project.taglib.xml file (as I use Facelets) looks like: <facelet-taglib> <namespace>http://my.project/jsf</namespace> <tag> <tag-name>dataTable</tag-name> <component> <component-type>my.component.dataTable</component-type> <renderer-type>my.renderkit.dataTable</renderer-type> </component> </tag> So as you can see, I have two classes in my project for my custom datatable: MyHtmlDataTable and MyDataTableRenderer. One of my idea is to modify the content of the <f:facet> directly in the doEncodeBegin() method of my renderer. This is working (in fact almost working), but I don't really think that's the better place to achieve my modification. What do you think? Technical information: JSF 1.2, Facelets, Richfaces 3.3.2, Java 1.6

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  • Using JSON Data to Populate a Google Map with Database Objects

    - by MikeH
    I'm revising this question after reading the resources mentioned in the original answers and working through implementing it. I'm using the google maps api to integrate a map into my Rails site. I have a markets model with the following columns: ID, name, address, lat, lng. On my markets/index view, I want to populate a map with all the markets in my markets table. I'm trying to output @markets as json data, and that's where I'm running into problems. I have the basic map displaying, but right now it's just a blank map. I'm following the tutorials very closely, but I can't get the markers to generate dynamically from the json. Any help is much appreciated! Here's my setup: Markets Controller: def index @markets = Market.filter_city(params[:filter]) respond_to do |format| format.html # index.html.erb format.json { render :json => @market} format.xml { render :xml => @market } end end Markets/index view: <head> <script type="text/javascript" src="http://www.google.com/jsapi?key=GOOGLE KEY REDACTED, BUT IT'S THERE" > </script> <script type="text/javascript"> var markets = <%= @markets.to_json %>; </script> <script type="text/javascript" charset="utf-8"> google.load("maps", "2.x"); google.load("jquery", "1.3.2"); </script> </head> <body> <div id="map" style="width:400px; height:300px;"></div> </body> Public/javascripts/application.js: function initialize() { if (GBrowserIsCompatible() && typeof markets != 'undefined') { var map = new GMap2(document.getElementById("map")); map.setCenter(new GLatLng(40.7371, -73.9903), 13); map.addControl(new GLargeMapControl()); function createMarker(latlng, market) { var marker = new GMarker(latlng); var html="<strong>"+market.name+"</strong><br />"+market.address; GEvent.addListener(marker,"click", function() { map.openInfoWindowHtml(latlng, html); }); return marker; } var bounds = new GLatLngBounds; for (var i = 0; i < markets.length; i++) { var latlng=new GLatLng(markets[i].lat,markets[i].lng) bounds.extend(latlng); map.addOverlay(createMarker(latlng, markets[i])); } } } window.onload=initialize; window.onunload=GUnload;

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  • Is there a programming language with be semantics close to English ?

    - by ivo s
    Most languages allow to 'tweek' to certain extend parts of the syntax (C++,C#) and/or semantics that you will be using in your code (Katahdin, lua). But I have not heard of a language that can just completely define how your code will look like. So isn't there some language which already exists that has such capabilities to override all syntax & define semantics ? Example of what I want to do is basically from the C# code below: foreach(Fruit fruit in Fruits) { if(fruit is Apple) { fruit.Price = fruit.Price/2; } } I want do be able to to write the above code in my perfect language like this: Check if any fruits are Macintosh apples and discount the price by 50%. The advantages that come to my mind looking from a coder's perspective in this "imaginary" language are: It's very clear what is going on (self descriptive) - it's plain English after all even kid would understand my program Hides all complexities which I have to write in C#. But why should I care to learn that if statements, arithmetic operators etc since there are already implemented The disadvantages that I see for a coder who will maintain this program are: Maybe you would express this program differently from me so you may not get all the information that I've expressed in my sentence Programs can be quite verbose and hard to debug but if possible to even proximate this type of syntax above maybe more people would start programming right? That would be amazing I think. I can go to work and just write an essay to draw a square on a winform like this: Create a form called MyGreetingForm. Draw a square with in the middle of MyGreetingFormwith a side of 100 points. In the middle of the square write "Hello! Click here to continue" in Arial font. In the above code the parser must basically guess that I want to use the unnamed square from the previous sentence, it'd be hard to write such a smart parser I guess, yet it's so simple what I want to do. If the user clicks on square in the middle of MyGreetingForm show MyMainForm. In the above code 'basically' the compiler must: 1)generate an event handler 2) check if there is any square in the middle of the form and if there is - 3) hide the form and show another form It looks very hard to do but it doesn't look impossible IMO to me at least approximate this (I can personally generate a parser to perform the 3 steps above np & it's basically the same that it has to do any way when you add even in c# a.MyEvent=+handler; so I don't see a problem here) so I'm thinking maybe somebody already did something like this ? Or is there some practical burden of complexity to create such a 'essay style' programming language which I can't see ? I mean what's the worse that can happen if the parser is not that good? - your program will crash so you have to re-word it:)

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  • jQuery: Giving each matched element an unique ID

    - by AnGafraidh
    I am writing an 'inline translator' application to be used with a cloud computing platform to extend non-supported languages. The majority of this uses jQuery to find the text value, replace it with the translation, then append the element with a span tag that has an unique ID, to be used elsewhere within the application. The problem arises however, when there are more than one element, say , that have the exact same value to be translated (matched elements). What happens in the function in question is that it puts all matched elements in the same span, taking the second, third, fourth, etc. out of their parent tags. My code is pretty much like this example: <script src='jquery-1.4.2.js'></script> <script> jQuery.noConflict(); var uniqueID='asdfjkl'; jQuery(window).ready(function() { var myQ1 = jQuery("input[id~=test1]"); myClone=myQ1.clone(); myClone.val('Replaced this button'); myQ1.replaceWith('<span id='+uniqueID+'></span>'); jQuery('#'+uniqueID).append(myClone); }); </script> <table> <tr><td> <input id='test1' type='button' value="I'm a button!"></input> &nbsp; <input id='test2' type='button' value="And so am I"></input> </tr></td> <tr><td> <input id='test1' type='button' value="I'm a button!"></input> </tr></td> </table> As a workaround, I've experimented with using a loop to create a class for each span, rising in increment until jQuery("input[id~=test1]").length, but I can't seem to get anything I do to work. Is there any way to give each matched element an unique ID? My fluency in jQuery is being put to the test! Thanks for any help in advance. Aaron

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  • What web platform is right for me?

    - by egervari
    I've been looking at web frameworks like Rails, Grails, etc. I'm used to doing applications in Spring Framework with Hibernate... and I want something more productive. One of the things I realized is that while some of the things in Grails is sexy, there are some serious problems with it. Grails' controllers: 1) are implemented awfully. They don't seem to be able to extend from super classes at runtime. I tried this to add base actions and helper methods, and this seems to cause grails to blow up. 2) are based on an obsolete request parameters model (rather than form backing objects, which are much nicer). 3) are hard to test. Command objects are treated totally differently... and it's actually MUCH harder to write the test than it is to write the controller code. 4) Command objects operate totally differently. They are pre-validated and bound, which causes a lot of inconsistencies than basic parameter model. 5) Command objects are not reusable, and it's a pain in the rear to reuse most of the stuff from the domain classes, like constraints and fields. This is TRIVIAL to do in basic Spring. Why the hell was it not trivial to do in Grails? 6) The scaffolding that is generated is pure crap. It doesn't generalize inserts and updates... and it actually copy/pastes a pile of code in two views: create.gsp and edit.gsp. The views themselves are gargantuan piles of doggie do-do. This is further compounded by the fact that it uses low-level parameters and not objects. Integration tests are 30x slower than a Spring integration test. It is disgusting. Some mocking tests are so hard to write and aren't guaranteed to work when it's deployed, that I think it discourages fast, tdd test cycles. Most things seem to screw up grails while it's running, like adding a taglib, or anything really. The server restart problem wasn't solved at all. I'm starting to think going with Spring/Hibernate/Java is the only way to go. While there is a pretty big cost at startup, I know it'll eventually smooth out. It sucks I can't use a language like Scala... because idiomatically, it is so incompatible with Hibernate. This app is also not a run-of-the-mill UI over a database. It's got some of that, but it's not going to be a slouch. I am deathly scared of Grails now because of how crap it is in the Controller layer. Suggestions on what I can do?

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  • Simple Form validation failing Backbone

    - by Corey Buchillon
    Im not exactly adept at coding so Im probably missing something, but my view here is failing to refuse submission when one or both of the fields are empty. I have a feeling something isnt connected right to my template for the row and the view of the form Form = Backbone.View.extend({ //form vie el: '.item-form', initialize: function(){ }, events: { 'click #additem': 'addModel' }, addModel: function(itemName, price){ // simple validation before adding to collection if (itemName !="" && price !="" ){ var item = new Item({ itemName: this.$("#item").val(), price: this.$("#price").val()}); items.add(item); $("#message").html("Please wait; the task is being added."); item.save(null, {success: function (item, response,options) { item.id= item.attributes._id.$id; item.attributes.id = item.attributes._id.$id; new ItemsView({collection: items}); $("#message").html(""); } }); this.$("#item").val(''); this.$("#price").val(''); } else { alert('Please fill in both fields'); } } }); and HTML <table class="itemTable"> <thead> <tr> <th>Item</th> <th>Price</th> <th></th> </tr> </thead> <tbody class="tableBody"> <script type="text/template" id="table-row"> <td><%= itemName %></td> <td><%= price %></td> <td><button class="complete">Complete</button> <button class="remove">Remove</button></td> </script> </tbody> </table> <form class="item-form"> <input type="text" name="item" id="item" placeholder="Item"/> <!-- goes to itemName in the template for the body --> <input type="text" name="price" id="price" placeholder="Price" /><!--goes to price in the template for the body --> <button type="button" id="additem">Add</button> </form> <div id="message"></div>

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  • backbone.js - Having multiple instances of the same view

    - by TrueWheel
    I am having problems having multiple instances in of the same view in different div elements. When I try to initialize them only the second of the two elements appear no matter what order I put them in. Here is the code for my view. var BodyShapeView = Backbone.View.extend({ thingiview: null, scene: null, renderer: null, model: null, mouseX: 0, mouseY: 0, events:{ 'click button#front' : 'front', 'click button#diag' : 'diag', 'click button#in' : 'zoomIn', 'click button#out' : 'zoomOut', 'click button#on' : 'rotateOn', 'click button#off' : 'rotateOff', 'click button#wireframeOn' : 'wireOn', 'click button#wireframeOff' : 'wireOff', 'click button#distance' : 'dijkstra' }, initialize: function(name){ _.bindAll(this, 'render', 'animate'); scene = new THREE.Scene(); camera = new THREE.PerspectiveCamera( 15, 400 / 700, 1, 4000 ); camera.position.z = 3; scene.add( camera ); camera.position.y = -5; var ambient = new THREE.AmbientLight( 0x202020 ); scene.add( ambient ); var directionalLight = new THREE.DirectionalLight( 0xffffff, 0.75 ); directionalLight.position.set( 0, 0, 1 ); scene.add( directionalLight ); var pointLight = new THREE.PointLight( 0xffffff, 5, 29 ); pointLight.position.set( 0, -25, 10 ); scene.add( pointLight ); var loader = new THREE.OBJLoader(); loader.load( "img/originalMeanModel.obj", function ( object ) { object.children[0].geometry.computeFaceNormals(); var geometry = object.children[0].geometry; console.log(geometry); THREE.GeometryUtils.center(geometry); geometry.dynamic = true; var material = new THREE.MeshLambertMaterial({color: 0xffffff, shading: THREE.FlatShading, vertexColors: THREE.VertexColors }); mesh = new THREE.Mesh(geometry, material); model = mesh; // model = object; scene.add( model ); } ); // RENDERER renderer = new THREE.WebGLRenderer(); renderer.setSize( 400, 700 ); $(this.el).find('.obj').append( renderer.domElement ); this.animate(); }, Here is how I create the instances var morphableBody = new BodyShapeView({ el: $("#morphable-body") }); var bodyShapeView = new BodyShapeView({ el: $("#mean-body") }); Any help would be really appreciated. Thanks in advance.

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  • C++ class member functions instantiated by traits

    - by Jive Dadson
    I am reluctant to say I can't figure this out, but I can't figure this out. I've googled and searched Stack Overflow, and come up empty. The abstract, and possibly overly vague form of the question is, how can I use the traits-pattern to instantiate non-virtual member functions? The question came up while modernizing a set of multivariate function optimizers that I wrote more than 10 years ago. The optimizers all operate by selecting a straight-line path through the parameter space away from the current best point (the "update"), then finding a better point on that line (the "line search"), then testing for the "done" condition, and if not done, iterating. There are different methods for doing the update, the line-search, and conceivably for the done test, and other things. Mix and match. Different update formulae require different state-variable data. For example, the LMQN update requires a vector, and the BFGS update requires a matrix. If evaluating gradients is cheap, the line-search should do so. If not, it should use function evaluations only. Some methods require more accurate line-searches than others. Those are just some examples. The original version instantiates several of the combinations by means of virtual functions. Some traits are selected by setting mode bits that are tested at runtime. Yuck. It would be trivial to define the traits with #define's and the member functions with #ifdef's and macros. But that's so twenty years ago. It bugs me that I cannot figure out a whiz-bang modern way. If there were only one trait that varied, I could use the curiously recurring template pattern. But I see no way to extend that to arbitrary combinations of traits. I tried doing it using boost::enable_if, etc.. The specialized state information was easy. I managed to get the functions done, but only by resorting to non-friend external functions that have the this-pointer as a parameter. I never even figured out how to make the functions friends, much less member functions. The compiler (VC++ 2008) always complained that things didn't match. I would yell, "SFINAE, you moron!" but the moron is probably me. Perhaps tag-dispatch is the key. I haven't gotten very deeply into that. Surely it's possible, right? If so, what is best practice?

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  • jQuery question from a person who can't javascript

    - by Evilalan
    So I'm trying to adapt this Dropdown menu on Joomla the styles work great as expected so I'll post the javascript includes on the head of my website: <script type='text/javascript' src='js/jquery.js'></script> <script type='text/javascript' src='js/dropdown.js'></script> <script type='text/javascript'> $(function() { $('.menu').droppy(); }); </script> <script type='text/javascript'> $(function() { $('.menu').droppy({speed: 100}); }); </script> ok I don't know why its is not working I'll post the dropdown.js should I post the jQuery too? it's really big! $.fn.droppy = function(options) { options = $.extend({speed: 250}, options || {}); this.each(function() { var root = this, zIndex = 1000; function getSubnav(ele) { if (ele.nodeName.toLowerCase() == 'li') { var subnav = $('> ul', ele); return subnav.length ? subnav[0] : null; } else { return ele; } } function getActuator(ele) { if (ele.nodeName.toLowerCase() == 'ul') { return $(ele).parents('li')[0]; } else { return ele; } } function hide() { var subnav = getSubnav(this); if (!subnav) return; $.data(subnav, 'cancelHide', false); setTimeout(function() { if (!$.data(subnav, 'cancelHide')) { $(subnav).slideUp(options.speed); } }, 500); } function show() { var subnav = getSubnav(this); if (!subnav) return; $.data(subnav, 'cancelHide', true); $(subnav).css({zIndex: zIndex++}).slideDown(options.speed); if (this.nodeName.toLowerCase() == 'ul') { var li = getActuator(this); $(li).addClass('hover'); $('> a', li).addClass('hover'); } } $('ul, li', this).hover(show, hide); $('li', this).hover( function() { $(this).addClass('hover'); $('> a', this).addClass('hover'); }, function() { $(this).removeClass('hover'); $('> a', this).removeClass('hover'); } ); }); }; My question here is: Why is it not working! I know that this is really complex (I don't anything about JavaScript) but if you help me I'll post a tutorial and edited files that will help a lot of people! By the way I've download jQuery from the original site so I don't think that this can be the problem! Thanks in advance!

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  • Passing custom info to mongrel_rails start

    - by whaka
    One thing I really don't understand is how I can pass custom start-up options to a mongrel instance. I see that a common approach is the use environment variables, but in my environment this is not going to work because my rails application serves many different clients. Much code is shared between clients, but there are also many differences which I implement by subclassing controllers and views to overload or extend existing features or introduce new ones. To make this all work, I simply add the paths to client specific modules the module load path ($:). In order to start the application for a particular client, I could now use an environment variable like say, TARGET=AMAZONE. Unfortunately, on some systems I'm running multiple mongrel clusters, each cluster serving a different client. Some of these systems run under Windows and to start mongrel I installed mongrel_services. Clearly, this makes my environment variable unsuitable. Passing this extra bit of data to the application is proving to be a real challenge. For a start, mongrel_rails service_install will reject any [custom] command line parameters that aren't documented. I'm not too concerned as installing the services using the install program is trivial. However, even if I manage to install mongrel_services such that when run it passes the custom command line option --target to mongrel_rails start, I get an error because mongrel_rails doesn't recognize the switch. So here were the things I looked at: Pass an extra parameter: mongrel_rails start --target XYZ ... use a config file and add target:XYZ, then do: mongrel_rails start -C x:\myapp\myconfig.yml modify the file: Ruby\lib\ruby\gems\1.8\gems\mongrel-1.1.5-x86-mswin32-60\lib\mongrel\command.rb Perhaps I can use the --script option, but all docs that I found on it were for Unix 1 and 2 simply don't work. I played with 4 but never managed it to do anything. So I had no choice but to go with 3. While it is relatively simple, I hate changing ruby library code. Particularly disappointing is that 2 doesn't work. I mean what is so unreasonable about adding other [custom] options in the config file? Actually I think this is a fundamental piece that is missing in rails. Somehow, the application should be able to register and access command line arguments it expects. If anybody has a good idea how to do this more elegantly using the current infrastructure, I have a chocolate fish to give away!!!

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  • XSL transformation of SVG adds namespace attribute to new tag

    - by Steve
    I have a SVG file that I want to extend by adding onclick handlers to edges and nodes. I also want to add a script tag referring to a JavaScript. The problem is that the script tag gets an empty namespace attribute added to it. I haven't found any information regarding this that I understand. Why does XSLT add an empty namespace? XSL file: <?xml version="1.0"?> <xsl:stylesheet version="1.0" xmlns:xsl="http://www.w3.org/1999/XSL/Transform" xmlns:svg="http://www.w3.org/2000/svg" xmlns:xlink="http://www.w3.org/1999/xlink"> <xsl:output method="xml" encoding="utf-8" /> <xsl:template match="/svg:svg"> <xsl:copy> <script type="text/ecmascript" xlink:href="base.js" /> <!-- this tag gets a namespace attr --> <xsl:apply-templates /> </xsl:copy> </xsl:template> <!-- Identity transform http://www.w3.org/TR/xslt#copying --> <xsl:template match="@*|node()"> <xsl:copy> <xsl:apply-templates select="@*|node()"/> </xsl:copy> </xsl:template> <!-- Check groups and add functions --> <xsl:template match="svg:g"> <xsl:copy> <xsl:if test="@class = 'node'"> <xsl:attribute name="onclick">node_clicked()</xsl:attribute> </xsl:if> <xsl:if test="@class = 'edge'"> <xsl:attribute name="onclick">edge_clicked()</xsl:attribute> </xsl:if> <xsl:apply-templates select="@*|node()" /> </xsl:copy> </xsl:template> </xsl:stylesheet>

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  • JPA - Can I create an Entity class, using an @DiscriminatorValue, that doesn't have its own table?

    - by DaveyDaveDave
    Hi - this is potentially a bit complex, so I'll do my best to describe my situation - it's also my first post here, so please forgive formatting mistakes, etc! I'm using JPA with joined inheritance and a database structure that looks like: ACTION --------- ACTION_ID ACTION_MAPPING_ID ACTION_TYPE DELIVERY_CHANNEL_ACTION -------------------------- ACTION_ID CHANNEL_ID OVERRIDE_ADDRESS_ACTION -------------------------- ACTION_ID (various fields specific to this action type) So, in plain English, I have multiple different types of action, all share an ACTION_MAPPING, which is referenced from the 'parent' ACTION table. DELIVERY_CHANNEL_ACTION and OVERRIDE_ADDRESS_ACTION both have extra, supplementary data of their own, and are mapped to ACTION with a FK. Real-world, I also have a 'suppress' action, but this doesn't have any supplementary data of its own, so it doesn't have a corresponding table - all it needs is an ACTION_MAPPING, which is stored in the ACTION table. Hopefully you're with me so far... I'm creating a new project from scratch, so am pretty flexible in what I can do, but obviously would like to get it right from the outset! My current implementation, which works, has three entities loosely defined as follows: @Entity @Table(name="ACTION") @Inheritance(strategy=InheritanceType.JOINED) @DiscriminatorValue("SUPPRESS") public class Action @Entity @Table(name="DELIVERY_CHANNEL_ACTION") @DiscriminatorValue("DELIVERY_CHANNEL") public class DeliveryChannelAction extends Action @Entity @Table(name="OVERRIDE_ADDRESS_ACTION") @DiscriminatorValue("OVERRIDE_ADDRESS") public class OverrideAddressAction extends Action That is - I have a concrete base class, Action, with a Joined inheritance strategy. DeliveryChannelAction and OverrideAddressAction both extend Action. What feels wrong here though, is that my Action class is the base class for these two actions, but also forms the concrete implementation for the suppress action. For the time being this works, but at some point more actions are likely to be added, and there's every chance that some of them will, like SUPPRESS, have no supplementary data, which will start to get difficult! So... what I would like to do, in the object model world, is to have Action be abstract, and create a SuppressAction class, which is empty apart from having a @DiscriminatorValue("SUPPRESS"). I've tried doing exactly what is described above, so, changing Action to: @Entity @Table(name="ACTION") @Inheritance(strategy=InheritanceType.JOINED) public abstract class Action and creating: @DiscriminatorValue("SUPPRESS") public class SuppressAction extends Action but no luck - it seems to work fine for DeliveryChannelAction and OverrideAddressAction, but when I try to create a SuppressAction and persist it, I get: java.lang.IllegalArgumentException: Object: com.mypackage.SuppressAction[actionId=null] is not a known entity type. at org.eclipse.persistence.internal.sessions.UnitOfWorkImpl.registerNewObjectForPersist(UnitOfWorkImpl.java:4147) at org.eclipse.persistence.internal.jpa.EntityManagerImpl.persist(EntityManagerImpl.java:368) at com.mypackage.test.util.EntityTestUtil.createSuppressAction(EntityTestUtil.java:672) at com.mypackage.entities.ActionTest.testCRUDAction(ActionTest.java:27) which I assume is down to the fact that SuppressAction isn't registered as an entity, but I don't know how I can do that, given that it doesn't have an associated table. Any pointers, either complete answers or hints for things to Google (I'm out of ideas!), most welcome :) EDIT: to correct my stacktrace.

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  • show tweets inside div from an asynchronous loop

    - by ak_47
    Am trying to laod tweets into a div after looping them from yahoo placemaker. They are loading on the div but the information shown by them is placemaker's last result. This is the code.. function getLocation(user, date, profile_img, text,url) { var templates = []; templates[0] = '<div><div></div><h2 class="firstHeading">'+user+'</h2><div>'+text+'</div><div><p><a href="' + url + '"target="_blank">'+url+'</a></p></div><p>Date Posted- '+date+'</p></div>'; templates[1] = '<table width="320" border="0"><tr><td class="user" colspan="2" rowspan="1">'+user+'</td></tr><tr><td width="45"><a href="'+profile_img+'"><img src="'+profile_img+'" width="55" height="50"/></a></td><td width="186">'+text+'<p><a href="' + url + '"target="_blank">'+url+'</a></p></td></tr></table><hr>'; templates[2] = '<div><div></div><h2 class="firstHeading">'+user+'</h2><div>'+text+'</div><div><p><a href="' + url + '"target="_blank">'+url+'</a></p></div><p>Date Posted- '+date+'</p></div>'; templates[3] = '<table width="320" border="0"><tr><td class="user" colspan="2" rowspan="1">'+user+'</td></tr><tr><td width="45"><a href="'+profile_img+'"><img src="'+profile_img+'" width="55" height="50"/></a></td><td width="186">'+text+'<p><a href="' + url + '"target="_blank">'+url+'</a></p></td></tr></table><hr>'; var geocoder = new google.maps.Geocoder(); Placemaker.getPlaces(text, function (o) { console.log(o); if (!$.isArray(o.match)) { var latitude = o.match.place.centroid.latitude; var longitude = o.match.place.centroid.longitude; var myLatLng = new google.maps.LatLng(latitude, longitude); var marker = new google.maps.Marker({ icon: profile_img, title: user, map: map, position: myLatLng }); var infowindow = new google.maps.InfoWindow({ content: templates[0].replace('user',user).replace('text',text).replace('url',url).replace('date',date) }); var $tweet = $(templates[1].replace('%user',user).replace(/%profile_img/g,profile_img).replace('%text',text).replace('%url',url)); $('#user-banner').css("visibility","visible");$('#news-banner').css("visibility","visible"); $('#news-tweets').css("overflow","scroll").append($tweet); function openInfoWindow() { infowindow.open(map, marker); } google.maps.event.addListener(marker, 'click', openInfoWindow); $tweet.find(".user").on('click', openInfoWindow); bounds.extend(myLatLng); } }); }

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  • Google Map + MarkerClusterer only takes place when map completely zooms out

    - by user415795
    The clustering works but somehow it only takes place at the maximum zoom out(the largest view with all nations), the moment I zoom in by 1 value, the clustering icon changes back to markers. I try with all kinds of values on the maxZoom and gridSize clusterer options with no help. Can someone please kindly advice. Thanks. <script language="javascript" type="text/javascript"> var markersArray = []; var mc = null; var markersArray = []; var mc = null; var map; var mapOptions; var geocoder; var infoWindow; var http_request = false; var lat = 0; var lng = 0; var startingZoom = 7; var lowestZoom = 1; // The lower the number, the more places can be seen on within the bounds. var highestZoom = 8; function mapLoad() { geocoder = new google.maps.Geocoder(); infoWindow = new google.maps.InfoWindow(); mapOptions = { zoomControl: true, zoom: 2, minZoom: lowestZoom, maxzoom: highestZoom, draggable: true, scrollwheel: true, disableDoubleClickZoom: true, mapTypeId: google.maps.MapTypeId.ROADMAP }; map = new google.maps.Map(document.getElementById('map'), mapOptions); } $(document).ready(function () { var searchUrl; var locations; // Place the user's current location marker on the map var Location = new google.maps.LatLng(1.340319, 103.743744); var Location2 = new google.maps.LatLng(1.322347, 103.757881); createMarker('1', Location, 'My Location', '', '', '', '/Images/home.png'); createMarker('1', Location2, 'My Location', '', '', '', '/Images/bb.png'); var bounds = new google.maps.LatLngBounds(); bounds.extend(gameLocation); map.fitBounds(bounds); }); // Create the marker with address information function createMarker(actId, point, address1, address2, town, postcode, icon) { var marker = new google.maps.Marker({ map: map, icon: icon, position: point, title: address1, animation: google.maps.Animation.DROP }); marker.metadata = { id: actId }; markersArray.push(marker); mc = new MarkerClusterer(map, markersArray); return marker; } </script>

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  • Help with simple frame, and graphics in Java

    - by Crystal
    For hw, I'm trying to create a "CustomButton" that has a frame and in that frame, I draw two triangles, and a square over it. It's supposed to give the user the effect of a button press once it is depressed. So for starters, I am trying to set up the beginning graphics, drawing two triangles, and a square. The problem I have is although I set my frame to 200, 200, and the triangles I have drawn I think to the correct ends of my frame size, when I run the program, I have to extend my window to make the whole artwork, my "CustomButton," viewable. Is that normal? Thanks. Code: import java.awt.*; import java.awt.event.*; import javax.swing.*; public class CustomButton { public static void main(String[] args) { EventQueue.invokeLater(new Runnable() { public void run() { CustomButtonFrame frame = new CustomButtonFrame(); frame.setDefaultCloseOperation(JFrame.EXIT_ON_CLOSE); frame.setVisible(true); } }); } } class CustomButtonFrame extends JFrame { // constructor for CustomButtonFrame public CustomButtonFrame() { setTitle("Custom Button"); setSize(DEFAULT_WIDTH, DEFAULT_HEIGHT); CustomButtonSetup buttonSetup = new CustomButtonSetup(); this.add(buttonSetup); } private static final int DEFAULT_WIDTH = 200; private static final int DEFAULT_HEIGHT = 200; } class CustomButtonSetup extends JComponent { public void paintComponent(Graphics g) { Graphics2D g2 = (Graphics2D) g; // first triangle coords int x[] = new int[TRIANGLE_SIDES]; int y[] = new int[TRIANGLE_SIDES]; x[0] = 0; y[0] = 0; x[1] = 200; y[1] = 0; x[2] = 0; y[2] = 200; Polygon firstTriangle = new Polygon(x, y, TRIANGLE_SIDES); // second triangle coords x[0] = 0; y[0] = 200; x[1] = 200; y[1] = 200; x[2] = 200; y[2] = 0; Polygon secondTriangle = new Polygon(x, y, TRIANGLE_SIDES); g2.drawPolygon(firstTriangle); g2.setColor(Color.WHITE); g2.fillPolygon(firstTriangle); g2.drawPolygon(secondTriangle); g2.setColor(Color.GRAY); g2.fillPolygon(secondTriangle); // draw rectangle 10 pixels off border g2.drawRect(10, 10, 180, 180); } public static final int TRIANGLE_SIDES = 3; }

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  • How do I mock/fake/replace/stub a base class at unit-test time in C#?

    - by MatthewMartin
    UPDATE: I've changed the wording of the question. Previously it was a yes/no question about if a base class could be changed at runtime. I may be working on mission impossible here, but I seem to be getting close. I want to extend a ASP.NET control, and I want my code to be unit testable. Also, I'd like to be able to fake behaviors of a real Label (namely things like ID generation, etc), which a real Label can't do in an nUnit host. Here a working example that makes assertions on something that depends on a real base class and something that doesn't-- in a more realistic unit test, the test would depend on both --i.e. an ID existing and some custom behavior. Anyhow the code says it better than I can: public class LabelWrapper : Label //Runtime //public class LabelWrapper : FakeLabel //Unit Test time { private readonly LabelLogic logic= new LabelLogic(); public override string Text { get { return logic.ProcessGetText(base.Text); } set { base.Text=logic.ProcessSetText(value); } } } //Ugh, now I have to test FakeLabelWrapper public class FakeLabelWrapper : FakeLabel //Unit Test time { private readonly LabelLogic logic= new LabelLogic(); public override string Text { get { return logic.ProcessGetText(base.Text); } set { base.Text=logic.ProcessSetText(value); } } } [TestFixture] public class UnitTest { [Test] public void Test() { //Wish this was LabelWrapper label = new LabelWrapper(new FakeBase()) LabelWrapper label = new LabelWrapper(); //FakeLabelWrapper label = new FakeLabelWrapper(); label.Text = "ToUpper"; Assert.AreEqual("TOUPPER",label.Text); StringWriter stringWriter = new StringWriter(); HtmlTextWriter writer = new HtmlTextWriter(stringWriter); label.RenderControl(writer); Assert.AreEqual(1,label.ID); Assert.AreEqual("<span>TOUPPER</span>", stringWriter.ToString()); } } public class FakeLabel { virtual public string Text { get; set; } public void RenderControl(TextWriter writer) { writer.Write("<span>" + Text + "</span>"); } } //System Under Test internal class LabelLogic { internal string ProcessGetText(string value) { return value.ToUpper(); } internal string ProcessSetText(string value) { return value.ToUpper(); } }

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  • A question about making a C# class persistant during a file load

    - by Adam
    Apologies for the indescriptive title, however it's the best I could think of for the moment. Basically, I've written a singleton class that loads files into a database. These files are typically large, and take hours to process. What I am looking for is to make a method where I can have this class running, and be able to call methods from within it, even if it's calling class is shut down. The singleton class is simple. It starts a thread that loads the file into the database, while having methods to report on the current status. In a nutshell it's al little like this: public sealed class BulkFileLoader { static BulkFileLoader instance = null; int currentCount = 0; BulkFileLoader() public static BulkFileLoader Instance { // Instanciate the instance class if necessary, and return it } public void Go() { // kick of 'ProcessFile' thread } public void GetCurrentCount() { return currentCount; } private void ProcessFile() { while (more rows in the import file) { // insert the row into the database currentCount++; } } } The idea is that you can get an instance of BulkFileLoader to execute, which will process a file to load, while at any time you can get realtime updates on the number of rows its done so far using the GetCurrentCount() method. This works fine, except the calling class needs to stay open the whole time for the processing to continue. As soon as I stop the calling class, the BulkFileLoader instance is removed, and it stops processing the file. What I am after is a solution where it will continue to run independently, regardless of what happens to the calling class. I then tried another approach. I created a simple console application that kicks off the BulkFileLoader, and then wrapped it around as a process. This fixes one problem, since now when I kick off the process, the file will continue to load even if I close the class that called the process. However, now the problem I have is that cannot get updates on the current count, since if I try and get the instance of BulkFileLoader (which, as mentioned before is a singleton), it creates a new instance, rather than returning the instance that is currently in the executing process. It would appear that singletons don't extend into the scope of other processes running on the machine. In the end, I want to be able to kick off the BulkFileLoader, and at any time be able to find out how many rows it's processed. However, that is even if I close the application I used to start it. Can anyone see a solution to my problem?

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  • PHP-OOP extending two classes?

    - by user1292810
    I am very beginner to OOP and now I am trying to write some PHP class to connect with FTP server. class ftpConnect { private $server; private $user; private $password; private $connection_id; private $connection_correct = false; public function __construct($server, $user = "anonymous", $password = "[email protected]") { $this->server = $server; $this->user = $user; $this->password = $password; $this->connection_id = ftp_connect($this->server); $this->connection_correct = ftp_login($this->connection_id, $this->user, $this->password); if ( (!$this->connection_id) || (!$this->connection_correct) ){ echo "Error! Couldn't connect to $this->server"; var_dump($this->connection_id); var_dump($this->connection_correct); return false; } else { echo "Successfully connected to $this->server, user: $this->user"; $this->connection_correct = true; return true; } } } I reckon that body of the class is insignificant at the moment. Main issue is that I have some problems with understanding OOP idea. I wanted to add sending emails every time, when the code is run. I have downloaded PHPMailer Class and extended my class with it: class ftpConnect extends PHPMailer {...} I have added some variables and methods and everything works as expected to that point. I thought: why not to add storing everything in database. Everytime user runs above code, proper information should be stored in database. I could edit my ftpConnect class and add database connecting to the constructor, and some other methods to updating tables. But database connecting and all that stuff could be used by other classes in the future, so it definitely should be implemented in seperate class. But my "main" ftpConnect class already extends one class and could not extend not a single one more. I have no idea how can I resolve this problem. Maybe my ftpConnect class is to complex and I should somehow divide it into couple smaller classes? Any help is much appreciated.

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  • Constructors from extended class in Java

    - by Crystal
    I'm having some trouble with a hw assignment. In one assignment, we had to create a Person class. Mine was: public class Person { String firstName; String lastName; String telephone; String email; public Person() { firstName = ""; lastName = ""; telephone = ""; email = ""; } public Person(String firstName) { this.firstName = firstName; } public Person(String firstName, String lastName, String telephone, String email) { this.firstName = firstName; this.lastName = lastName; this.telephone = telephone; this.email = email; } public String getFirstName() { return firstName; } public void setFirstName(String firstName) { this.firstName = firstName; } public String getLastName() { return lastName; } public void setLastName(String lastName) { this.lastName = lastName; } public String getTelephone() { return telephone; } public void setTelephone(String telephone) { this.telephone = telephone; } public String getEmail() { return email; } public void setEmail(String email) { this.email = email; } public boolean equals(Object otherObject) { // a quick test to see if the objects are identical if (this == otherObject) { return true; } // must return false if the explicit parameter is null if (otherObject == null) { return false; } if (!(otherObject instanceof Person)) { return false; } Person other = (Person) otherObject; return firstName.equals(other.firstName) && lastName.equals(other.lastName) && telephone.equals(other.telephone) && email.equals(other.email); } public int hashCode() { return 7 * firstName.hashCode() + 11 * lastName.hashCode() + 13 * telephone.hashCode() + 15 * email.hashCode(); } public String toString() { return getClass().getName() + "[firstName = " + firstName + '\n' + "lastName = " + lastName + '\n' + "telephone = " + telephone + '\n' + "email = " + email + "]"; } } Now we have to extend that class and use that class in our constructor. The function protoype is: public CarLoan(Person client, double vehiclePrice, double downPayment, double salesTax, double interestRate, CAR_LOAN_TERMS length) I'm confused on how I use the Person constructor from the superclass. I cannot necessarily do super(client); in my constructor which is what the book did with some primitive types in their example. Not sure what the correct thing to do is... Any thoughts? Thanks!

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  • Apple Airport Express, Extreme and Time Capsules, BT Home Hub, Wireless Extenders confusion

    - by Jamie Hartnoll
    I post quite frequently in Stack Overflow, but use Superuser less frequently. Mainly as I don't change hardware often and rarely have software issues! I live in a small stone cottage, and have an office in a separate building across a yard. I have a BT Homehub which is located in the cottage and a series of Ethernet cables running across the yard to the office. This is fine for my wired stuff. My main office computers are PCs running Windows 7 Ultimate, and one on Win7 Home, all working fine. I also have an old laptop on Win XP which works fine wirelessly in the house for those evenings in front of the TV catching up on a bit of work. I also have an iPhone and an iPad. Recently, I have been trying to get WiFi in the office so I can use Adobe Shadow (or whatever it now is!) to improve mobile web development efficiency using my iPhone and iPad, so I bought this: http://www.ebuyer.com/393462-zyxel-wre2205-500mbps-powerline-wireless-n300-range-extender-wre2205-gb0101f Thinking that would be lovely just plugged into the socket by the door in the office, extending the perimeter of the WiFi from my Homehub. I can't get it to work properly! If I plug a laptop into its ethernet port I can get it to connect to the Homehub and give me a kinda of wired, wireless extender. If, however, I plug the ethernet port into my home hub, it then seems to extend the network, but only my iOs devices work, and all my wired stuff stops working, and seems to create an infinite loop where windows connects to my homehob, and then rather to the internet, it then connects back to the extender thing. Anyway... in the meantime, I took a fatal trip to the Apple Store, where I purchased an Airport Express... solely for the purpose of hooking my iOs devices up as wireless music players in the house. I knew it had WiFi, but didn't want to use that part as an extender, I didn't think it would work on a Homehub anyway. It doesn't work on a Homehub! I now have a new wireless network in the house, which, when anything connects to it cannot connect to the Internet, so it works ONLY as a wireless music player. I then borrowed some Powerline Adaptors from someone and realised that this whole thing was getting totally out of control! It seems all the technology is out there but it's so complicated to get the right series of devices. To further add to the confusion, I wouldn't mind a network hard drive. I bought one that broke and lost everything, so now we're on to looking at the Apple Time Capsules. So my question is... IF... I buy an Apple Time Capsule, can I: Hook that up to my Homehub, leaving the homehub connected to the Internet so my Hub phones still work, then disable wireless on the homehub Link up my Airport Express to the Time Capsule PROPERLY so it will connect to the Internet Do the above with an Apple TV box should I buy one in future Use the Time Capsule as a network hard drive to store video and music that can be viewed/listened to via my iOS devices/Apple TV/Aiport Express anywhere even with my main PC off (this currently stores all this data) Hope that the IOS devices like the WiFi from the TimeCapsule better than the Homehub and work without extension, or buy another Airport Express to get WiFI in the office. Or... should I buy an Airport Extreme and use a USB hard drive for the network drive?

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  • Thoughts on home NAS server

    - by user826955
    I currently have a NAS with a 2x2TB HDD 1x16GB SSD layout on a mini-itx atom board. The NAS is in a Lian Li PC-Q07 case. On this system I was running freebsd 8 with a gmirror raid 1 setup, which was enough for my needs. So far I was using the NAS for: Fileserver with AFP protocol (only mac clients used) SVN server hosting all my source trees of my projects JIRA (performance was okay-ish) Timemachine backup for the macs The power consumption was about 38W, although I did not put HDDs asleep when unused (I think this is not possible in a raid setup). I liked the NAS because: the performance was good through gigabit LAN (enough for my needs) power consumption was good its a pretty small case and fits in one of my cupboards I disliked the NAS a bit because: it was a bit noisy, the Q07 case vibrated a good amount because of the HDDs. I switched the NAS off every evening I do not have a real backup of the data on the NAS, only the internal raid 1 as safety. I really dont want to loose my source trees under no circumstances, so I would really be sleeping better if I knew I had regular backups somewhere. Recently, the board seemed to have died, I can't boot anymore. Thus, I was thinking about a redesign of my NAS (I still have to find out what parts are broken, I probably need to replace the mainboard and SSD. HDDs seem to be okay). First of all, I was wondering what other users have as backup for their NAS? Are you actually using a second NAS, and regularly copying over the data to have it safe? Or is there any better solution to this? I was thinking about getting a cheap NAS like the synology DS112j with only one disk, and use rsync or something similar to regularly copy data over to the second NAS (wake the second NAS upon start, shut it down after copy). Although this approach seems somewhat weird, It would have the benefit (?) that I could use a single disk instead of raid in the main NAS, and put the disk asleep when idle, and have the NAS running 24/7 with low energy consumption (I found no way to do this with a gmirror setup). Is there any recommended backup solution for a small NAS? Then I was thinking about a different raid setup. Since I have to buy a new mainboard as well as SSD, I might as well switch over to a i3 board with more ram, and also switch to ZFS. I am not familar with ZFS, I've never used it, but I read and hear much about it. Would it be viable to set up a ZFS storage with only 2 disks? Can I easily extend this storage with more disks, once I choose to add some? I could maybe get a new case like the Fractal Design Array R2 which has more 3,5" slots. I could as well get another 2 disks, but I would prefer sticking with the existing 2 for enegery/heat/noise reasons. Should I go for a ZFS storage or stick to my gmirror setup? I would also like to keep freebsd as operating system, and also I dont need any web gui or something (that is, I dont need/want to use FreeNAS or Openfiler etc). Does anyone maybe have a sample setup in use so I can compare energy consumption/noise/software setup? Any guidance towards the NAS of my dreams (silent, low energy, safe w/ backups) much appreciated.

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  • Publishing an Excel spreadsheet using Microsoft SBS 2008 to a web page that is viewable by mobile ph

    - by Dave Heath
    I am getting well out of my “superuser” depth here and would love some support. At work we have an Excel workbook (*.xls format circa Office 2003) which maintains our “engineers” timesheet. This handles what events we are doing across the year and how many “work units” it is. As far as a workbook goes, it is fairly simple with just a few =SUM(range) cells and some linked across sheets (12 sheets, one for each month) It is stored on a server, in a folder that provides “management” with full access and “engineers” with read-only access. The workbook itself is read-only for “engineers” and full access for “management”. I think these permissions are controlled through Active Directory. The workbook is protected with a password, assumingly to allow “management” to edit it even if they are working at a terminal logged in as an “engineer”. This protection prevents “engineers” from going to certain cells to see formulae and therefore editing them. The workbook has a macro which saves and closes it ten minutes after opening. This is to stop other “management” from being locked out by any one person who has logged in with editing privileges. I hope this is making sense to someone... :S Now then, we have Microsoft Small Business Server 2008. We have a shiny new web-based login for when we are offsite so we can get to Exchange webmail and our internal site (which uses Sharepoint 3.0). “Management” would like to be able to publish this timesheet automatically after changes (they don’t want to have to do anything different to what they are currently doing) so that using an iPhone “engineers” can check on it while out of the office. I am currently having a look at “Excel Services” for Office 2007 on TechNet but I am not sure if I am running down the right garden path at the moment. < EDIT This seems to suggest that I have to have Sharepoint Server 2007, with no mention of Sharepoint 3.0... ... "MOSS builds on WSS by adding both core features as well as end user web parts" - Wikipedia entry for Microsoft Office SharePoint Server (MOSS) this is not good news... "...and using the ASP.NET APIs, web parts can be written to extend the functionality of WSS." Wikipedia entry for Windows Sharepoint Services. Could this bring back what I need? Is this good news? Do I need to start learning ASP.NET? This link here implies that we need MOSS to do what I want and the bosses say we aint' getting it. http://serverfault.com/questions/20198/what-is-some-cool-things-you-can-do-with-sharepoint-2007/22128#22128 Back to the drawing board. < /EDIT Please could someone suggest some “further reading” for me to help point me in the right direction or to put me back on the right track. Many thanks. I will try to keep this up to date with how I get on.

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  • Determining the required depth and specifications for a server cabinet

    - by Bingu Bingme
    I'm trying to understand the considerations ("why") that go into determining the specifications ("what") for a rackmount server cabinet, in order to determine what sort of rack I should purchase for my home use. Since this is for home use, I won't be following certain best practices (eg. hot/cold aisle, not even air conditioning) and may be willing to sacrifice in various areas in order to reduce cost and footprint - but please advise if there are safety concerns or other considerations to note. The most basic specs for a server cabinet are the dimensions (external width x external depth x usable height). Width: commonly 600mm or 800mm (if the use case requires extra clearance around the sides, such as if there is lots of cabling). In my case and most common cases, I'm going to stick with 600mm. Height: Select a sufficiently tall rack to fit my equipment. But how much may I stuff into it? Eg, if there is a 15U rack, can I really populate it with 15U of servers, or should I leave 1U at top and bottom for air circulation? Depth: Racks commonly have external depth of 600mm (network equipment), 800mm, 1000mm, or even longer. I'm trying to see how to fit into the 800mm depth. With reference to http://www.server-racks.com/rack-mount-depth.html, I'm hoping to have the front and rear posts mounted ~ 28.5" (72cm) apart, which would leave only 8cm for front space and rear space. How much rear space (from rear posts to back of rack) do I really need? I won't use cable management arms, so can I mount a 72cm depth server since the power, KVM, network cables won't take up much depth? My most important equipment are all < 60cm depth (4U chassis) and should comfortably fit within the 800mm cabinet. The rest of the equipment are very old 1U servers that range from 65-72cm depth. I might still want to make further use of them, or I might discard them since they are so old. Even if the 72cm servers cannot be powered on in an 800mm rack, I should be able to use them as 1U shelves. But, what server depth can I expect to be able to operate? Or am I forced to upgrade to 1000mm depth racks in order to use any servers deeper than 60cm? With reference to best practices for HP racks, some other specs and installation considerations: There aren't any minimum recommendations for clearance on the sides of the rack. It is recommended to leave 48" front clearance. The 48" front clearance is based on 32" chassis depth, 13" to extend the rack rails and mate the inner/outer rails, and 3" for movement. If I don't use such rails (eg, use shelves instead), it should be sufficient to leave front clearance of chassis depth + 3". It is recommended to leave 30" rear clearance "to provide space for servicing the rack". I'm planning to back the rack into a corner of the room, and wheel it slightly out when I need to access the rear. If the wheeling plan is ok, I still need to know how much rear clearance is required for air circulation and ventilation purposes. Castor wheels and stabilising feet. Since I'm backing the rack into a corner of the room, I'll only be able to set the stabilising feet on the front corners. Thoughts on safety? The rack that I'm considering has front glass doors with side ventilation slits and fully perforated rear doors. I'm hoping this will be a good balance between temperature and noise (only ventilation slits facing out the front, while the rear is facing the walls). Or is the sound of high-rpm fans going to escape through the front slits anyway and destroy my sanity?

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  • Multi-tenant ASP.NET MVC – Introduction

    - by zowens
    I’ve read a few different blogs that talk about multi-tenancy and how to resolve some of the issues surrounding multi-tenancy. What I’ve come to realize is that these implementations overcomplicate the issues and give only a muddy implementation! I’ve seen some really illogical code out there. I have recently been building a multi-tenancy framework for internal use at eagleenvision.net. Through this process, I’ve realized a few different techniques to make building multi-tenant applications actually quite easy. I will be posting a few different entries over the issue and my personal implementation. In this first post, I will discuss what multi-tenancy means and how my implementation will be structured.   So what’s the problem? Here’s the deal. Multi-tenancy is basically a technique of code-reuse of web application code. A multi-tenant application is an application that runs a single instance for multiple clients. Here the “client” is different URL bindings on IIS using ASP.NET MVC. The problem with different instances of the, essentially, same application is that you have to spin up different instances of ASP.NET. As the number of running instances of ASP.NET grows, so does the memory footprint of IIS. Stack Exchange shifted its architecture to multi-tenancy March. As the blog post explains, multi-tenancy saves cost in terms of memory utilization and physical disc storage. If you use the same code base for many applications, multi-tenancy just makes sense. You’ll reduce the amount of work it takes to synchronize the site implementations and you’ll thank your lucky stars later for choosing to use one application for multiple sites. Multi-tenancy allows the freedom of extensibility while relying on some pre-built code.   You’d think this would be simple. I have actually seen a real lack of reference material on the subject in terms of ASP.NET MVC. This is somewhat surprising given the number of users of ASP.NET MVC. However, I will certainly fill the void ;). Implementing a multi-tenant application takes a little thinking. It’s not straight-forward because the possibilities of implementation are endless. I have yet to see a great implementation of a multi-tenant MVC application. The only one that comes close to what I have in mind is Rob Ashton’s implementation (all the entries are listed on this page). There’s some really nasty code in there… something I’d really like to avoid. He has also written a library (MvcEx) that attempts to aid multi-tenant development. This code is even worse, in my honest opinion. Once I start seeing Reflection.Emit, I have to assume the worst :) In all seriousness, if his implementation makes sense to you, use it! It’s a fine implementation that should be given a look. At least look at the code. I will reference MvcEx going forward as a comparison to my implementation. I will explain why my approach differs from MvcEx and how it is better or worse (hopefully better).   Core Goals of my Multi-Tenant Implementation The first, and foremost, goal is to use Inversion of Control containers to my advantage. As you will see throughout this series, I pass around containers quite frequently and rely on their use heavily. I will be using StructureMap in my implementation. However, you could probably use your favorite IoC tool instead. <RANT> However, please don’t be stupid and abstract your IoC tool. Each IoC is powerful and by abstracting the capabilities, you’re doing yourself a real disservice. Who in the world swaps out IoC tools…? No one!</RANT> (It had to be said.) I will outline some of the goodness of StructureMap as we go along. This is really an invaluable tool in my tool belt and simple to use in my multi-tenant implementation. The second core goal is to represent a tenant as easily as possible. Just as a dependency container will be a first-class citizen, so will a tenant. This allows us to easily extend and use tenants. This will also allow different ways of “plugging in” tenants into your application. In my implementation, there will be a single dependency container for a single tenant. This will enable isolation of the dependencies of the tenant. The third goal is to use composition as a means to delegate “core” functions out to the tenant. More on this later.   Features In MvcExt, “Modules” are a code element of the infrastructure. I have simplified this concept and have named this “Features”. A feature is a simple element of an application. Controllers can be specified to have a feature and actions can have “sub features”. Each tenant can select features it needs and the other features will be hidden to the tenant’s users. My implementation doesn’t require something to be a feature. A controller can be common to all tenants. For example, (as you will see) I have a “Content” controller that will return the CSS, Javascript and Images for a tenant. This is common logic to all tenants and shouldn’t be hidden or considered a “feature”; Content is a core component.   Up next My next post will be all about the code. I will reveal some of the foundation to the way I do multi-tenancy. I will have posts dedicated to Foundation, Controllers, Views, Caching, Content and how to setup the tenants. Each post will be in-depth about the issues and implementation details, while adhering to my core goals outlined in this post. As always, comment with questions of DM me on twitter or send me an email.

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