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  • Manage SQL Server Connectivity through Windows Azure Virtual Machines Remote PowerShell

    - by SQLOS Team
    Manage SQL Server Connectivity through Windows Azure Virtual Machines Remote PowerShell Blog This blog post comes from Khalid Mouss, Senior Program Manager in Microsoft SQL Server. Overview The goal of this blog is to demonstrate how we can automate through PowerShell connecting multiple SQL Server deployments in Windows Azure Virtual Machines. We would configure TCP port that we would open (and close) though Windows firewall from a remote PowerShell session to the Virtual Machine (VM). This will demonstrate how to take the advantage of the remote PowerShell support in Windows Azure Virtual Machines to automate the steps required to connect SQL Server in the same cloud service and in different cloud services.  Scenario 1: VMs connected through the same Cloud Service 2 Virtual machines configured in the same cloud service. Both VMs running different SQL Server instances on them. Both VMs configured with remote PowerShell turned on to be able to run PS and other commands directly into them remotely in order to re-configure them to allow incoming SQL connections from a remote VM or on premise machine(s). Note: RDP (Remote Desktop Protocol) is kept configured in both VMs by default to be able to remote connect to them and check the connections to SQL instances for demo purposes only; but not actually required. Step 1 – Provision VMs and Configure Ports   Provision VM1; named DemoVM1 as follows (see examples screenshots below if using the portal):   Provision VM2 (DemoVM2) with PowerShell Remoting enabled and connected to DemoVM1 above (see examples screenshots below if using the portal): After provisioning of the 2 VMs above, here is the default port configurations for example: Step2 – Verify / Confirm the TCP port used by the database Engine By the default, the port will be configured to be 1433 – this can be changed to a different port number if desired.   1. RDP to each of the VMs created below – this will also ensure the VMs complete SysPrep(ing) and complete configuration 2. Go to SQL Server Configuration Manager -> SQL Server Network Configuration -> Protocols for <SQL instance> -> TCP/IP - > IP Addresses   3. Confirm the port number used by SQL Server Engine; in this case 1433 4. Update from Windows Authentication to Mixed mode   5.       Restart SQL Server service for the change to take effect 6.       Repeat steps 3., 4., and 5. For the second VM: DemoVM2 Step 3 – Remote Powershell to DemoVM1 Enter-PSSession -ComputerName condemo.cloudapp.net -Port 61503 -Credential <username> -UseSSL -SessionOption (New-PSSessionOption -SkipCACheck -SkipCNCheck) Your will then be prompted to enter the password. Step 4 – Open 1433 port in the Windows firewall netsh advfirewall firewall add rule name="DemoVM1Port" dir=in localport=1433 protocol=TCP action=allow Output: netsh advfirewall firewall show rule name=DemoVM1Port Rule Name:                            DemoVM1Port ---------------------------------------------------------------------- Enabled:                              Yes Direction:                            In Profiles:                             Domain,Private,Public Grouping:                             LocalIP:                              Any RemoteIP:                             Any Protocol:                             TCP LocalPort:                            1433 RemotePort:                           Any Edge traversal:                       No Action:                               Allow Ok. Step 5 – Now connect from DemoVM2 to DB instance in DemoVM1 Step 6 – Close port 1433 in the Windows firewall netsh advfirewall firewall delete rule name=DemoVM1Port Output: Deleted 1 rule(s). Ok. netsh advfirewall firewall show  rule name=DemoVM1Port No rules match the specified criteria.   Step 7 – Try to connect from DemoVM2 to DB Instance in DemoVM1  Because port 1433 has been closed (in step 6) in the Windows Firewall in VM1 machine, we can longer connect from VM3 remotely to VM1. Scenario 2: VMs provisioned in different Cloud Services 2 Virtual machines configured in different cloud services. Both VMs running different SQL Server instances on them. Both VMs configured with remote PowerShell turned on to be able to run PS and other commands directly into them remotely in order to re-configure them to allow incoming SQL connections from a remote VM or on on-premise machine(s). Note: RDP (Remote Desktop Protocol) is kept configured in both VMs by default to be able to remote connect to them and check the connections to SQL instances for demo purposes only; but not actually needed. Step 1 – Provision new VM3 Provision VM3; named DemoVM3 as follows (see examples screenshots below if using the portal): After provisioning is complete, here is the default port configurations: Step 2 – Add public port to VM1 connect to from VM3’s DB instance Since VM3 and VM1 are not connected in the same cloud service, we will need to specify the full DNS address while connecting between the machines which includes the public port. We shall add a public port 57000 in this case that is linked to private port 1433 which will be used later to connect to the DB instance. Step 3 – Remote Powershell to DemoVM1 Enter-PSSession -ComputerName condemo.cloudapp.net -Port 61503 -Credential <UserName> -UseSSL -SessionOption (New-PSSessionOption -SkipCACheck -SkipCNCheck) You will then be prompted to enter the password.   Step 4 – Open 1433 port in the Windows firewall netsh advfirewall firewall add rule name="DemoVM1Port" dir=in localport=1433 protocol=TCP action=allow Output: Ok. netsh advfirewall firewall show rule name=DemoVM1Port Rule Name:                            DemoVM1Port ---------------------------------------------------------------------- Enabled:                              Yes Direction:                            In Profiles:                             Domain,Private,Public Grouping:                             LocalIP:                              Any RemoteIP:                             Any Protocol:                             TCP LocalPort:                            1433 RemotePort:                           Any Edge traversal:                       No Action:                               Allow Ok.   Step 5 – Now connect from DemoVM3 to DB instance in DemoVM1 RDP into VM3, launch SSM and Connect to VM1’s DB instance as follows. You must specify the full server name using the DNS address and public port number configured above. Step 6 – Close port 1433 in the Windows firewall netsh advfirewall firewall delete rule name=DemoVM1Port   Output: Deleted 1 rule(s). Ok. netsh advfirewall firewall show  rule name=DemoVM1Port No rules match the specified criteria.  Step 7 – Try to connect from DemoVM2 to DB Instance in DemoVM1  Because port 1433 has been closed (in step 6) in the Windows Firewall in VM1 machine, we can no longer connect from VM3 remotely to VM1. Conclusion Through the new support for remote PowerShell in Windows Azure Virtual Machines, one can script and automate many Virtual Machine and SQL management tasks. In this blog, we have demonstrated, how to start a remote PowerShell session, re-configure Virtual Machine firewall to allow (or disallow) SQL Server connections. References SQL Server in Windows Azure Virtual Machines   Originally posted at http://blogs.msdn.com/b/sqlosteam/

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  • HTML5-MVC application using VS2010 SP1

    - by nmarun
    This is my first attempt at creating HTML5 pages. VS 2010 allows working with HTML5 now (you just need to make a small change after installing SP1). So my Razor view is now a HTML5 page. I call this application - 5Commerce – (an over-simplified) HTML5 ECommerce site. So here’s the flow of the application: home page renders user enters first and last name, chooses a product and the quantity can enter additional instructions for the order place the order user is then taken to another page showing the order details Off to the details. This is what my page looks in Google Chrome 10 beta (or later) soon after it renders. Here are some of the things to observe on this. Look a little closer and you’ll see a border around the first name textbox – this is ‘autofocus’ in action. I’ve set the autofocus attribute on this textbox. So as soon as the page loads, this control gets focus. 1: <input type="text" autofocus id="firstName" class="inputWidth" data_minlength="" 2: data_maxlength="" placeholder="first name" /> See a partially grayed out ‘last name’ text in the second textbox. This is set using a placeholder attribute (see above). It gets wiped out on-focus and improves the UI visuals in general. The quantity textbox is actually a numerical-only textbox. 1: <input type="number" id="quantity" data_mincount="" class="inputWidth" /> The last line is for additional instructions. This looks like a label but it’s content is editable. Just adding the ‘contenteditable’ attribute to the span allow the user to edit the text inside. 1: <span contenteditable id="additionalInstructions" data_texttype="" class="editableContent">select text and edit </span> All of the above is just plain HTML (no lurking javascript acting in here). Makes it real clean and simple. Going more into the HTML, I see that the _Layout.cshtml already is using some HTML5 content. I created my project before installing SP1, so that was the reason for my surprise. 1: <!DOCTYPE html> This is the doctype declaration in HTML5 and this is supported even by IE6 (just take my word on IE6 now, don’t go install it to test it, especially when MS is doing an IE6 countdown). That’s just amazing and extremely easy to read remember and talk about a few less bytes on every call! I modified the rest of my _Layout.cshtml to the below: 1: <!DOCTYPE html> 2: <html> 3: <head> 4: <title>5Commerce - HTML 5 Ecommerce site</title> 5: <link href="@Url.Content("~/Content/Site.css")" rel="stylesheet" type="text/css" /> 6: <script src="@Url.Content("~/Scripts/jquery-1.4.4.min.js")" type="text/javascript"></script> 7: <script src="@Url.Content("~/Scripts/CustomScripts.js")" type="text/javascript"></script> 8: <script type="text/javascript"> 9: $(document).ready(function () { 10: WireupEvents(); 11: }); 12:</script> 13:  14: </head> 15:  16: <body role="document" class="bodybackground"> 17: <header role="heading"> 18: <h2>5Commerce - HTML 5 Ecommerce site!</h2> 19: </header> 20: <section id="mainForm"> 21: @RenderBody() 22: </section> 23: <footer id="page_footer" role="siteBaseInfo"> 24: <p>&copy; 2011 5Commerce Inc!</p> 25: </footer> 26: </body> 27: </html> I’m sure you’re seeing some of the new tags here. To give a brief intro about them: <header>, <footer>: Marks the header/footer region of a page or section. <section>: A logical grouping of content role attribute: Identifies the responsibility of an element. This attribute can be used by screen readers and can also be filtered through jQuery. SP1 also allows for some intellisense in HTML5. You see the other types of input fields – email, date, datetime, month, url and there are others as well. So once my page loads, i.e., ‘on document ready’, I’m wiring up the events following the principles of unobtrusive javascript. In the snippet below, I’m controlling the behavior of the input controls for specific events. 1: $("#productList").bind('change blur', function () { 2: IsSelectedProductValid(); 3: }); 4:  5: $("#quantity").bind('blur', function () { 6: IsQuantityValid(); 7: }); 8:  9: $("#placeOrderButton").click( 10: function () { 11: if (IsPageValid()) { 12: LoadProducts(); 13: } 14: }); This enables some client-side validation to occur before the data is sent to the server. These validation constraints are obtained through a JSON call to the WCF service and are set to the ‘data_’ attributes of the input controls. Have a look at the ‘GetValidators()’ function below: 1: function GetValidators() { 2: // the post to your webservice or page 3: $.ajax({ 4: type: "GET", //GET or POST or PUT or DELETE verb 5: url: "http://localhost:14805/OrderService.svc/GetValidators", // Location of the service 6: data: "{}", //Data sent to server 7: contentType: "application/json; charset=utf-8", // content type sent to server 8: dataType: "json", //Expected data format from server 9: processdata: true, //True or False 10: success: function (result) {//On Successfull service call 11: if (result.length > 0) { 12: for (i = 0; i < result.length; i++) { 13: if (result[i].PropertyName == "FirstName") { 14: if (result[i].MinLength > 0) { 15: $("#firstName").attr("data_minLength", result[i].MinLength); 16: } 17: if (result[i].MaxLength > 0) { 18: $("#firstName").attr("data_maxLength", result[i].MaxLength); 19: } 20: } 21: else if (result[i].PropertyName == "LastName") { 22: if (result[i].MinLength > 0) { 23: $("#lastName").attr("data_minLength", result[i].MinLength); 24: } 25: if (result[i].MaxLength > 0) { 26: $("#lastName").attr("data_maxLength", result[i].MaxLength); 27: } 28: } 29: else if (result[i].PropertyName == "Quantity") { 30: if (result[i].MinCount > 0) { 31: $("#quantity").attr("data_minCount", result[i].MinCount); 32: } 33: } 34: else if (result[i].PropertyName == "AdditionalInstructions") { 35: if (result[i].TextType.length > 0) { 36: $("#additionalInstructions").attr("data_textType", result[i].TextType); 37: } 38: } 39: } 40: } 41: }, 42: error: function (result) {// When Service call fails 43: alert('Service call failed: ' + result.status + ' ' + result.statusText); 44: } 45: }); 46:  47: //.... 48: } Just before the GetValidators() function runs and sets the validation constraints, this is what the html looks like (seen through the Dev tools of Chrome): After the function executes, you see the values in the ‘data_’  attributes. As and when we enter valid data into these fields, the error messages disappear, since the validation is bound to the blur event of the control. There you see… no error messages (well, the catch here is that once you enter THAT name, all errors disappear automatically). Clicking on ‘Place Order!’ runs the SaveOrder function. You can see the JSON for the order object that is getting constructed and passed to the WCF Service. 1: function SaveOrder() { 2: var addlInstructionsDefaultText = "select text and edit"; 3: var addlInstructions = $("span:first").text(); 4: if(addlInstructions == addlInstructionsDefaultText) 5: { 6: addlInstructions = ''; 7: } 8: var orderJson = { 9: AdditionalInstructions: addlInstructions, 10: Customer: { 11: FirstName: $("#firstName").val(), 12: LastName: $("#lastName").val() 13: }, 14: OrderedProduct: { 15: Id: $("#productList").val(), 16: Quantity: $("#quantity").val() 17: } 18: }; 19:  20: // the post to your webservice or page 21: $.ajax({ 22: type: "POST", //GET or POST or PUT or DELETE verb 23: url: "http://localhost:14805/OrderService.svc/SaveOrder", // Location of the service 24: data: JSON.stringify(orderJson), //Data sent to server 25: contentType: "application/json; charset=utf-8", // content type sent to server 26: dataType: "json", //Expected data format from server 27: processdata: false, //True or False 28: success: function (result) {//On Successfull service call 29: window.location.href = "http://localhost:14805/home/ShowOrderDetail/" + result; 30: }, 31: error: function (request, error) {// When Service call fails 32: alert('Service call failed: ' + request.status + ' ' + request.statusText); 33: } 34: }); 35: } The service saves this order into an XML file and returns the order id (a guid). On success, I redirect to the ShowOrderDetail action method passing the guid. This page will show all the details of the order. Although the back-end weightlifting is done by WCF, I did not show any of that plumbing-work as I wanted to concentrate more on the HTML5 and its associates. However, you can see it all in the source here. I do have one issue with HTML5 and this is an existing issue with HTML4 as well. If you see the snippet above where I’ve declared a textbox for first name, you’ll see the autofocus attribute just dangling by itself. It doesn’t follow the xml syntax of ‘key="value"’ allowing users to continue writing badly-formatted html even in the new version. You’ll see the same issue with the ‘contenteditable’ attribute as well. The work-around is that you can do ‘autofocus=”true”’ and it’ll work fine plus make it well-formatted. But unless the standards enforce this, there will be people (me included) who’ll get by, by just typing the bare minimum! Hoping this will get fixed in the coming version-updates. Source code here. Verdict: I think it’s time for us to embrace the new HTML5. Thank you HTML4 and Welcome HTML5.

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  • Metro: Dynamically Switching Templates with a WinJS ListView

    - by Stephen.Walther
    Imagine that you want to display a list of products using the WinJS ListView control. Imagine, furthermore, that you want to use different templates to display different products. In particular, when a product is on sale, you want to display the product using a special “On Sale” template. In this blog entry, I explain how you can switch templates dynamically when displaying items with a ListView control. In other words, you learn how to use more than one template when displaying items with a ListView control. Creating the Data Source Let’s start by creating the data source for the ListView. Nothing special here – our data source is a list of products. Two of the products, Oranges and Apples, are on sale. (function () { "use strict"; var products = new WinJS.Binding.List([ { name: "Milk", price: 2.44 }, { name: "Oranges", price: 1.99, onSale: true }, { name: "Wine", price: 8.55 }, { name: "Apples", price: 2.44, onSale: true }, { name: "Steak", price: 1.99 }, { name: "Eggs", price: 2.44 }, { name: "Mushrooms", price: 1.99 }, { name: "Yogurt", price: 2.44 }, { name: "Soup", price: 1.99 }, { name: "Cereal", price: 2.44 }, { name: "Pepsi", price: 1.99 } ]); WinJS.Namespace.define("ListViewDemos", { products: products }); })(); The file above is saved with the name products.js and referenced by the default.html page described below. Declaring the Templates and ListView Control Next, we need to declare the ListView control and the two Template controls which we will use to display template items. The markup below appears in the default.html file: <!-- Templates --> <div id="productItemTemplate" data-win-control="WinJS.Binding.Template"> <div class="product"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> </div> <div id="productOnSaleTemplate" data-win-control="WinJS.Binding.Template"> <div class="product onSale"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> </div> <!-- ListView --> <div id="productsListView" data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource: ListViewDemos.products.dataSource, layout: { type: WinJS.UI.ListLayout } }"> </div> In the markup above, two Template controls are declared. The first template is used when rendering a normal product and the second template is used when rendering a product which is on sale. The second template, unlike the first template, includes the text “(On Sale!)”. The ListView control is bound to the data source which we created in the previous section. The ListView itemDataSource property is set to the value ListViewDemos.products.dataSource. Notice that we do not set the ListView itemTemplate property. We set this property in the default.js file. Switching Between Templates All of the magic happens in the default.js file. The default.js file contains the JavaScript code used to switch templates dynamically. Here’s the entire contents of the default.js file: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { WinJS.UI.processAll().then(function () { var productsListView = document.getElementById("productsListView"); productsListView.winControl.itemTemplate = itemTemplateFunction; });; } }; function itemTemplateFunction(itemPromise) { return itemPromise.then(function (item) { // Select either normal product template or on sale template var itemTemplate = document.getElementById("productItemTemplate"); if (item.data.onSale) { itemTemplate = document.getElementById("productOnSaleTemplate"); }; // Render selected template to DIV container var container = document.createElement("div"); itemTemplate.winControl.render(item.data, container); return container; }); } app.start(); })(); In the code above, a function is assigned to the ListView itemTemplate property with the following line of code: productsListView.winControl.itemTemplate = itemTemplateFunction;   The itemTemplateFunction returns a DOM element which is used for the template item. Depending on the value of the product onSale property, the DOM element is generated from either the productItemTemplate or the productOnSaleTemplate template. Using Binding Converters instead of Multiple Templates In the previous sections, I explained how you can use different templates to render normal products and on sale products. There is an alternative approach to displaying different markup for normal products and on sale products. Instead of creating two templates, you can create a single template which contains separate DIV elements for a normal product and an on sale product. The following default.html file contains a single item template and a ListView control bound to the template. <!-- Template --> <div id="productItemTemplate" data-win-control="WinJS.Binding.Template"> <div class="product" data-win-bind="style.display: onSale ListViewDemos.displayNormalProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> <div class="product onSale" data-win-bind="style.display: onSale ListViewDemos.displayOnSaleProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> </div> <!-- ListView --> <div id="productsListView" data-win-control="WinJS.UI.ListView" data-win-options="{ itemDataSource: ListViewDemos.products.dataSource, itemTemplate: select('#productItemTemplate'), layout: { type: WinJS.UI.ListLayout } }"> </div> The first DIV element is used to render a normal product: <div class="product" data-win-bind="style.display: onSale ListViewDemos.displayNormalProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> </div> The second DIV element is used to render an “on sale” product: <div class="product onSale" data-win-bind="style.display: onSale ListViewDemos.displayOnSaleProduct"> <span data-win-bind="innerText:name"></span> <span data-win-bind="innerText:price"></span> (On Sale!) </div> Notice that both templates include a data-win-bind attribute. These data-win-bind attributes are used to show the “normal” template when a product is not on sale and show the “on sale” template when a product is on sale. These attributes set the Cascading Style Sheet display attribute to either “none” or “block”. The data-win-bind attributes take advantage of binding converters. The binding converters are defined in the default.js file: (function () { "use strict"; var app = WinJS.Application; app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { WinJS.UI.processAll(); } }; WinJS.Namespace.define("ListViewDemos", { displayNormalProduct: WinJS.Binding.converter(function (onSale) { return onSale ? "none" : "block"; }), displayOnSaleProduct: WinJS.Binding.converter(function (onSale) { return onSale ? "block" : "none"; }) }); app.start(); })(); The ListViewDemos.displayNormalProduct binding converter converts the value true or false to the value “none” or “block”. The ListViewDemos.displayOnSaleProduct binding converter does the opposite; it converts the value true or false to the value “block” or “none” (Sadly, you cannot simply place a NOT operator before the onSale property in the binding expression – you need to create both converters). The end result is that you can display different markup depending on the value of the product onSale property. Either the contents of the first or second DIV element are displayed: Summary In this blog entry, I’ve explored two approaches to displaying different markup in a ListView depending on the value of a data item property. The bulk of this blog entry was devoted to explaining how you can assign a function to the ListView itemTemplate property which returns different templates. We created both a productItemTemplate and productOnSaleTemplate and displayed both templates with the same ListView control. We also discussed how you can create a single template and display different markup by using binding converters. The binding converters are used to set a DIV element’s display property to either “none” or “block”. We created a binding converter which displays normal products and a binding converter which displays “on sale” products.

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  • Taking web sites offline for demonstration

    While working in software development in general, and in web development for a couple of customers it is quite common that it is necessary to provide a test bed where the client is able to get an image, or better said, a feeling for the visions and ideas you are talking about. Usually here at IOS Indian Ocean Software Ltd. we set up a demo web site on one of our staging servers, and provide credentials to the customer to access and review our progress and work ad hoc. This gives us the highest flexibility on both sides, as the test bed is simply online and available 24/7. We can update the structure, the UI and data at any time, and the client is able to view it as it suits best for her/him. Limited or lack of online connectivity But what is going to happen when your client is not capable to be online - no matter for what reasons; here are some more obvious ones: No internet connection (permanently or temporarily) Expensive connection, ie. mobile data package, stay at a hotel, etc. Presentation devices at an exhibition, ie. using tablets or iPads Being abroad for a certain time, and only occasionally online No network coverage, especially on mobile Bad infrastructure, like ie. in Third World countries Providing a catalogue on CD or USB pen drive Anyway, it doesn't matter really. We should be able to provide a solution for the circumstances of our customers. Presentation during an exhibition Recently, we had the following request from a customer: Is it possible to let us have a desktop version of ResortWork.co.uk that we can use for demo purposes at the forthcoming Ski Shows? It would allow us to let stand visitors browse the sites on an iPad to view jobs and training directory course listings. Yes, sure we can do that. Eventually, you might think why don't they simply use 3G enabled iPads for that purpose? As stated above, there might be several reasons for that - low coverage, expensive data packages, etc. Anyway, it is not a question on how to circumvent the request but to deliver a solution to that. Possible solutions... or not? We already did offline websites earlier, and even established complete mirrors of one or two web sites on our systems. There are actually several possibilities to handle this kind of request, and it mainly depends on the system or device where the offline site should be available on. Here, it is clearly expressed that we have to address this on an Apple iPad, well actually, I think that they'd like to use multiple devices during their exhibitions. Following is an overview of possible solutions depending on the technology or device in use, and how it can be done: Replication of source files and database The above mentioned web site is running on ASP.NET, IIS and SQL Server. In case that a laptop or slate runs a Windows OS, the easiest way would be to take a snapshot of the source files and database, and transfer them as local installation to those Windows machines. This approach would be fully operational on the local machine. Saving pages for offline usage This is actually a quite tedious job but still practicable for small web sites Tool based approach to 'harvest' the web site There quite some tools in the wild that could handle this job, namely wget, httrack, web copier, etc. Screenshots bundled as PDF document Not really... ;-) Creating screencast or video Simply navigate through your website and record your desktop session. Actually, we are using this kind of approach to track down difficult problems in order to see and understand exactly what the user was doing to cause an error. Of course, this list isn't complete and I'd love to get more of your ideas in the comments section below the article. Preparations for offline browsing The original website is dynamically and data-driven by ASP.NET, and looks like this: As we have to put the result onto iPads we are going to choose the tool-based approach to 'download' the whole web site for offline usage. Again, depending on the complexity of your web site you might have to check which of the applications produces the best results for you. My usual choice is to use wget but in this case, we run into problems related to the rewriting of hyperlinks. As a consequence of that we opted for using HTTrack. HTTrack comes in different flavours, like console application but also as either GUI (WinHTTrack on Windows) or Web client (WebHTTrack on Linux/Unix/BSD). Here's a brief description taken from the original website about HTTrack: HTTrack is a free (GPL, libre/free software) and easy-to-use offline browser utility. It allows you to download a World Wide Web site from the Internet to a local directory, building recursively all directories, getting HTML, images, and other files from the server to your computer. HTTrack arranges the original site's relative link-structure. Simply open a page of the "mirrored" website in your browser, and you can browse the site from link to link, as if you were viewing it online. And there is an extensive documentation for all options and switches online. General recommendation is to go through the HTTrack Users Guide By Fred Cohen. It covers all the initial steps you need to get up and running. Be aware that it will take quite some time to get all the necessary resources down to your machine. Actually, for our customer we run the tool directly on their web server to avoid unnecessary traffic and bandwidth. After a couple of runs and some additional fine-tuning - explicit inclusion or exclusion of various external linked web sites - we finally had a more or less complete offline version available. A very handsome feature of HTTrack is the error/warning log after completing the download. It contains some detailed information about errors that appeared on the pages and the links within the pages that have been processed. Error: "Bad Request" (400) at link www.resortwork.co.uk/job-details_Ski_hire:tech_or_mgr_or_driver_37854.aspx (from www.resortwork.co.uk/Jobs_A_to_Z.aspx)Error: "Not Found" (404) at link www.247recruit.net/images/applynow.png (from www.247recruit.net/css/global.css)Error: "Not Found" (404) at link www.247recruit.net/activate.html (from www.247recruit.net/247recruit_tefl_jobs_network.html) In our situation, we took the records of HTTP 400/404 errors and passed them to the web development department. Improvements are to be expected soon. ;-) Quality assurance on the full-featured desktop Unfortunately, the generated output of HTTrack was still incomplete but luckily there were only images missing. Being directly on the web server we simply copied the missing images from the original source folder into our offline version. After that, we created an archive and transferred the file securely to our local workspace for further review and checks. From that point on, it wasn't necessary to get any more files from the original web server, and we could focus ourselves completely on the process of browsing and navigating through the offline version to isolate visual differences and functional problems. As said, the original web site runs on ASP.NET Web Forms and uses Postback calls for interaction like search, pagination and partly for navigation. This is the main field of improving the offline experience. Of course, same as for standard web development it is advised to test with various browsers, and strangely we discovered that the offline version looked pretty good on Firefox, Chrome and Safari, but not in Internet Explorer. A quick look at the HTML source shed some light on this, and there are conditional CSS inclusions based on the user agent. HTTrack is not acting as Internet Explorer and so we didn't have the necessary overrides for this browser. Not problematic after all in our case, but you might have to pay attention to this and get the IE-specific files explicitly. And while having a view at the source code, we also found out that HTTrack actually modifies the generated HTML output. In several occasions we discovered that <div> elements were converted into <table> constructs for no obvious reason; even nested structures. Search 'e'nd destroy - sed (or Notepad++) to the rescue During our intensive root cause analysis for a couple of HTML/CSS problems that needed some extra attention it is very helpful to be familiar with any editor that allows search and replace over multiple files like, ie. sed - stream editor for filtering and transforming text on Linux or my personal favourite Notepad++ on Windows. This allowed us to quickly fix a lot of anchors with onclick attributes and Javascript code that was addressed to ASP.NET files instead of their generated HTML counterparts, like so: grep -lr -e '.aspx' * | xargs sed -e 's/.aspx/.html?/g' The additional question mark after the HTML extension helps to separate the query string from the actual target and solved all our missing hyperlinks very fast. The same can be done in Notepad++ on Windows, too. Just use the 'Replace in files' feature and you are settled. Especially, in combination with Regular Expressions (regex). Landscape of browsers Okay, after several runs of HTML/CSS code analysis, searching and replacing some strings in a pool of more than 4.000 files, we finally had a very good match of an offline browsing experience in Firefox and Chrome on Linux. Next, we transferred that modified set of files to a Windows 8 machine for review on Firefox, Chrome and Internet Explorer 7 to 10, and a Mac mini running Mac OS X 10.7 to check the output on Safari and again on Chrome. Besides IE, for reasons already mentioned above, the results were identical. And last but not least it was about to check web site on tablets. Please continue to read on the following articles: Taking web sites offline for demonstration on Galaxy Tablet Taking web sites offline for demonstration on iPad

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  • ASP.NET MVC 2 from Scratch &ndash; Part 1 Listing Data from Database

    - by Max
    Part 1 - Listing Data from Database: Let us now learn ASP.NET MVC 2 from Scratch by actually developing a front end website for the Chinook database, which is an alternative to the traditional Northwind database. You can get the Chinook database from here. As always the best way to learn something is by working on it and doing something. The Chinook database has the following schema, a quick look will help us implementing the application in a efficient way. Let us first implement a grid view table with the list of Employees with some details, this table also has the Details, Edit and Delete buttons on it to do some operations. This is series of post will concentrate on creating a simple CRUD front end for Chinook DB using ASP.NET MVC 2. In this post, we will look at listing all the possible Employees in the database in a tabular format, from which, we can then edit and delete them as required. In this post, we will concentrate on setting up our environment and then just designing a page to show a tabular information from the database. We need to first setup the SQL Server database, you can download the required version and then set it up in your localhost. Then we need to add the LINQ to SQL Classes required for us to enable interaction with our database. Now after you do the above step, just use your Server Explorer in VS 2010 to actually navigate to the database, expand the tables node and then drag drop all the tables onto the Object Relational Designer space and you go you will have the tables visualized as classes. As simple as that. Now for the purpose of displaying the data from Employee in a table, we will show only the EmployeeID, Firstname and lastname. So let us create a class to hold this information. So let us add a new class called EmployeeList to the ViewModels. We will send this data model to the View and this can be displayed in the page. public class EmployeeList { public int EmployeeID { get; set; } public string Firstname { get; set; } public string Lastname { get; set; } public EmployeeList(int empID, string fname, string lname) { this.EmployeeID = empID; this.Firstname = fname; this.Lastname = lname; } } Ok now we have got the backend ready. Let us now look at the front end view now. We will first create a master called Site.Master and reuse it across the site. The Site.Master content will be <%@ Master Language="C#" AutoEventWireup="true" CodeBehind="Site.Master.cs" Inherits="ChinookMvcSample.Views.Shared.Site" %>   <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head id="Head1" runat="server"> <title></title> <style type="text/css"> html { background-color: gray; } .content { width: 880px; position: relative; background-color: #ffffff; min-width: 880px; min-height: 800px; float: inherit; text-align: justify; } </style> <script src="../../Scripts/jquery-1.4.1.min.js" type="text/javascript"></script> <asp:ContentPlaceHolder ID="head" runat="server"> </asp:ContentPlaceHolder> </head> <body> <center> <h1> My Website</h1> <div class="content"> <asp:ContentPlaceHolder ID="body" runat="server"> </asp:ContentPlaceHolder> </div> </center> </body> </html> The backend Site.Master.cs does not contain anything. In the actual Index.aspx view, we add the code to simply iterate through the collection of EmployeeList that was sent to the View via the Controller. So in the top of the Index.aspx view, we have this inherits which says Inherits="System.Web.Mvc.ViewPage<IEnumerable<ChinookMvcSample.ViewModels.EmployeeList>>" In this above line, we dictate that the page is consuming a IEnumerable collection of EmployeeList. So once we specify this and compile the project. Then in our Index.aspx page, we can consume the EmployeeList object and access all its methods and properties. <table class="styled" cellpadding="3" border="0" cellspacing="0"> <tr> <th colspan="3"> </th> <th> First Name </th> <th> Last Name </th> </tr> <% foreach (var item in Model) { %> <tr> <td align="center"> <%: Html.ActionLink("Edit", "Edit", new { id = item.EmployeeID }, new { id = "links" })%> </td> <td align="center"> <%: Html.ActionLink("Details", "Details", new { id = item.EmployeeID }, new { id = "links" })%> </td> <td align="center"> <%: Html.ActionLink("Delete", "Delete", new { id = item.EmployeeID }, new { id = "links" })%> </td> <td> <%: item.Firstname %> </td> <td> <%: item.Lastname %> </td> </tr> <% } %> <tr> <td colspan="5"> <%: Html.ActionLink("Create New", "Create") %> </td> </tr> </table> The Html.ActionLink is a Html Helper to a create a hyperlink in the page, in the one we have used, the first parameter is the text that is to be used for the hyperlink, second one is the action name, third one is the parameter to be passed, last one is the attributes to be added while the hyperlink is rendered in the page. Here we are adding the id=”links” to the hyperlinks that is created in the page. In the index.aspx page, we add some jQuery stuff add alternate row colours and highlight colours for rows on mouse over. Now the Controller that handles the requests and directs the request to the right view. For the index view, the controller would be public ActionResult Index() { //var Employees = from e in data.Employees select new EmployeeList(e.EmployeeId,e.FirstName,e.LastName); //return View(Employees.ToList()); return View(_data.Employees.Select(p => new EmployeeList(p.EmployeeId, p.FirstName, p.LastName))); } Let us also write a unit test using NUnit for the above, just testing EmployeeController’s Index. DataClasses1DataContext _data; public EmployeeControllerTest() { _data = new DataClasses1DataContext("Data Source=(local);Initial Catalog=Chinook;Integrated Security=True"); }   [Test] public void TestEmployeeIndex() { var e = new EmployeeController(_data); var result = e.Index() as ViewResult; var employeeList = result.ViewData.Model; Assert.IsNotNull(employeeList, "Result is null."); } In the first EmployeeControllerTest constructor, we set the data context to be used while running the tests. And then in the actual test, We just ensure that the View results returned by Index is not null. Here is the zip of the entire solution files until this point. Let me know if you have any doubts or clarifications. Cheers! Have a nice day.

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  • Context Sensitive JTable

    - by Geertjan
    Here's a plain old JTable on the NetBeans Platform. Whenever the toolbar button is clicked, information about the currently selected row is displayed in the status bar: Normally, the above would be achieved in NetBeans Platform applications via Nodes publishing their underlying business object when the selection changes. In this case, there are no Nodes at all. There's only a JTable and a DefaultTableModel, i.e., all pure Java Swing. So, how does it work? To follow the logic, it makes sense to create the example yourself, starting with the Stock object: public class Stock {     String name;     String desc;     public Stock() {     }     public Stock(String name, String desc) {         this.name = name;         this.desc = desc;     }     public String getDesc() {         return desc;     }     public String getName() {         return name;     }     public void setDesc(String desc) {         this.desc = desc;     }     public void setName(String name) {         this.name = name;     } } Next, create a new Window Component via the wizard and then rewrite the constructor as follows: public final class MyWindowTopComponent extends TopComponent {     private final InstanceContent ic = new InstanceContent();     public MyWindowTopComponent() {         initComponents();         //Statically create a few stocks,         //in reality these would come from a data source         //of some kind:         List<Stock> list = new ArrayList();         list.add(new Stock("AMZN", "Amazon"));         list.add(new Stock("BOUT", "About.com"));         list.add(new Stock("Something", "Something.com"));         //Create a JTable, passing the List above         //to a DefaultTableModel:         final JTable table = new JTable(StockTableModel (list));         //Whenever the mouse is clicked on the table,         //somehow construct a new Stock object //(or get it from the List above) and publish it:         table.addMouseListener(new MouseAdapter() {             @Override             public void mousePressed(MouseEvent e) {                 int selectedColumn = table.getSelectedColumn();                 int selectedRow = table.getSelectedRow();                 Stock s = new Stock();                 if (selectedColumn == 0) {                     s.setName(table.getModel().getValueAt(selectedRow, 0).toString());                     s.setDesc(table.getModel().getValueAt(selectedRow, 1).toString());                 } else {                     s.setName(table.getModel().getValueAt(selectedRow, 1).toString());                     s.setDesc(table.getModel().getValueAt(selectedRow, 0).toString());                 }                 ic.set(Collections.singleton(s), null);             }         });         JScrollPane scrollPane = new JScrollPane(table);         add(scrollPane, BorderLayout.CENTER);         //Put the dynamic InstanceContent into the Lookup:         associateLookup(new AbstractLookup(ic));     }     private DefaultTableModel StockTableModel (List<Stock> stockList) {         DefaultTableModel stockTableModel = new DefaultTableModel() {             @Override             public boolean isCellEditable(int row, int column) {                 return false;             }         };         Object[] columnNames = new Object[2];         columnNames[0] = "Symbol";         columnNames[1] = "Name";         stockTableModel.setColumnIdentifiers(columnNames);         Object[] rows = new Object[2];         ListIterator<Stock> stockListIterator = stockList.listIterator();         while (stockListIterator.hasNext()) {             Stock nextStock = stockListIterator.next();             rows[0] = nextStock.getName();             rows[1] = nextStock.getDesc();             stockTableModel.addRow(rows);         }         return stockTableModel;     }     ...     ...     ... And now, since you're publishing a new Stock object whenever the user clicks in the table, you can create loosely coupled Actions, like this: @ActionID(category = "Edit", id = "org.my.ui.ShowStockAction") @ActionRegistration(iconBase = "org/my/ui/Datasource.gif", displayName = "#CTL_ShowStockAction") @ActionReferences({     @ActionReference(path = "Menu/File", position = 1300),     @ActionReference(path = "Toolbars/File", position = 300) }) @Messages("CTL_ShowStockAction=Show Stock") public final class ShowStockAction implements ActionListener {     private final Stock context;     public ShowStockAction(Stock context) {         this.context = context;     }     @Override     public void actionPerformed(ActionEvent ev) {         StatusDisplayer.getDefault().setStatusText(context.getName() + " / " + context.getDesc());     } }

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  • Refactoring Part 1 : Intuitive Investments

    - by Wes McClure
    Fear, it’s what turns maintaining applications into a nightmare.  Technology moves on, teams move on, someone is left to operate the application, what was green is now perceived brown.  Eventually the business will evolve and changes will need to be made.  The approach to those changes often dictates the long term viability of the application.  Fear of change, lack of passion and a lack of interest in understanding the domain often leads to a paranoia to do anything that doesn’t involve duct tape and bailing twine.  Don’t get me wrong, those have a place in the short term viability of a project but they don’t have a place in the long term.  Add to it “us versus them” in regards to the original team and those that maintain it, internal politics and other factors and you have a recipe for disaster.  This results in code that quickly becomes unmanageable.  Even the most clever of designs will eventually become sub optimal and debt will amount that exponentially makes changes difficult.  This is where refactoring comes in, and it’s something I’m very passionate about.  Refactoring is about improving the process whereby we make change, it’s an exponential investment in the process of change. Without it we will incur exponential complexity that halts productivity. Investments, especially in the long term, require intuition and reflection.  How can we tackle new development effectively via evolving the original design and paying off debt that has been incurred? The longer we wait to ask and answer this question, the more it will cost us.  Small requests don’t warrant big changes, but realizing when changes now will pay off in the long term, and especially in the short term, is valuable. I have done my fair share of maintaining applications and continuously refactoring as needed, but recently I’ve begun work on a project that hasn’t had much debt, if any, paid down in years.  This is the first in a series of blog posts to try to capture the process which is largely driven by intuition of smaller refactorings from other projects. Signs that refactoring could help: Testability How can decreasing test time not pay dividends? One of the first things I found was that a very important piece often takes 30+ minutes to test.  I can only imagine how much time this has cost historically, but more importantly the time it might cost in the coming weeks: I estimate at least 10-20 hours per person!  This is simply unacceptable for almost any situation.  As it turns out, about 6 hours of working with this part of the application and I was able to cut the time down to under 30 seconds!  In less than the lost time of one week, I was able to fix the problem for all future weeks! If we can’t test fast then we can’t change fast, nor with confidence. Code is used by end users and it’s also used by developers, consider your own needs in terms of the code base.  Adding logic to enable/disable features during testing can help decouple parts of an application and lead to massive improvements.  What exactly is so wrong about test code in real code?  Often, these become features for operators and sometimes end users.  If you cannot run an integration test within a test runner in your IDE, it’s time to refactor. Readability Are variables named meaningfully via a ubiquitous language? Is the code segmented functionally or behaviorally so as to minimize the complexity of any one area? Are aspects properly segmented to avoid confusion (security, logging, transactions, translations, dependency management etc) Is the code declarative (what) or imperative (how)?  What matters, not how.  LINQ is a great abstraction of the what, not how, of collection manipulation.  The Reactive framework is a great example of the what, not how, of managing streams of data. Are constants abstracted and named, or are they just inline? Do people constantly bitch about the code/design? If the code is hard to understand, it will be hard to change with confidence.  It’s a large undertaking if the original designers didn’t pay much attention to readability and as such will never be done to “completion.”  Make sure not to go over board, instead use this as you change an application, not in lieu of changes (like with testability). Complexity Simplicity will never be achieved, it’s highly subjective.  That said, a lot of code can be significantly simplified, tidy it up as you go.  Refactoring will often converge upon a simplification step after enough time, keep an eye out for this. Understandability In the process of changing code, one often gains a better understanding of it.  Refactoring code is a good way to learn how it works.  However, it’s usually best in combination with other reasons, in effect killing two birds with one stone.  Often this is done when readability is poor, in which case understandability is usually poor as well.  In the large undertaking we are making with this legacy application, we will be replacing it.  Therefore, understanding all of its features is important and this refactoring technique will come in very handy. Unused code How can deleting things not help? This is a freebie in refactoring, it’s very easy to detect with modern tools, especially in statically typed languages.  We have VCS for a reason, if in doubt, delete it out (ok that was cheesy)! If you don’t know where to start when refactoring, this is an excellent starting point! Duplication Do not pray and sacrifice to the anti-duplication gods, there are excellent examples where consolidated code is a horrible idea, usually with divergent domains.  That said, mediocre developers live by copy/paste.  Other times features converge and aren’t combined.  Tools for finding similar code are great in the example of copy/paste problems.  Knowledge of the domain helps identify convergent concepts that often lead to convergent solutions and will give intuition for where to look for conceptual repetition. 80/20 and the Boy Scouts It’s often said that 80% of the time 20% of the application is used most.  These tend to be the parts that are changed.  There are also parts of the code where 80% of the time is spent changing 20% (probably for all the refactoring smells above).  I focus on these areas any time I make a change and follow the philosophy of the Boy Scout in cleaning up more than I messed up.  If I spend 2 hours changing an application, in the 20%, I’ll always spend at least 15 minutes cleaning it or nearby areas. This gives a huge productivity edge on developers that don’t. Ironically after a short period of time the 20% shrinks enough that we don’t have to spend 80% of our time there and can move on to other areas.   Refactoring is highly subjective, never attempt to refactor to completion!  Learn to be comfortable with leaving one part of the application in a better state than others.  It’s an evolution, not a revolution.  These are some simple areas to look into when making changes and can help get one started in the process.  I’ve often found that refactoring is a convergent process towards simplicity that sometimes spans a few hours but often can lead to massive simplifications over the timespan of weeks and months of regular development.

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  • Solving Big Problems with Oracle R Enterprise, Part II

    - by dbayard
    Part II – Solving Big Problems with Oracle R Enterprise In the first post in this series (see https://blogs.oracle.com/R/entry/solving_big_problems_with_oracle), we showed how you can use R to perform historical rate of return calculations against investment data sourced from a spreadsheet.  We demonstrated the calculations against sample data for a small set of accounts.  While this worked fine, in the real-world the problem is much bigger because the amount of data is much bigger.  So much bigger that our approach in the previous post won’t scale to meet the real-world needs. From our previous post, here are the challenges we need to conquer: The actual data that needs to be used lives in a database, not in a spreadsheet The actual data is much, much bigger- too big to fit into the normal R memory space and too big to want to move across the network The overall process needs to run fast- much faster than a single processor The actual data needs to be kept secured- another reason to not want to move it from the database and across the network And the process of calculating the IRR needs to be integrated together with other database ETL activities, so that IRR’s can be calculated as part of the data warehouse refresh processes In this post, we will show how we moved from sample data environment to working with full-scale data.  This post is based on actual work we did for a financial services customer during a recent proof-of-concept. Getting started with the Database At this point, we have some sample data and our IRR function.  We were at a similar point in our customer proof-of-concept exercise- we had sample data but we did not have the full customer data yet.  So our database was empty.  But, this was easily rectified by leveraging the transparency features of Oracle R Enterprise (see https://blogs.oracle.com/R/entry/analyzing_big_data_using_the).  The following code shows how we took our sample data SimpleMWRRData and easily turned it into a new Oracle database table called IRR_DATA via ore.create().  The code also shows how we can access the database table IRR_DATA as if it was a normal R data.frame named IRR_DATA. If we go to sql*plus, we can also check out our new IRR_DATA table: At this point, we now have our sample data loaded in the database as a normal Oracle table called IRR_DATA.  So, we now proceeded to test our R function working with database data. As our first test, we retrieved the data from a single account from the IRR_DATA table, pull it into local R memory, then call our IRR function.  This worked.  No SQL coding required! Going from Crawling to Walking Now that we have shown using our R code with database-resident data for a single account, we wanted to experiment with doing this for multiple accounts.  In other words, we wanted to implement the split-apply-combine technique we discussed in our first post in this series.  Fortunately, Oracle R Enterprise provides a very scalable way to do this with a function called ore.groupApply().  You can read more about ore.groupApply() here: https://blogs.oracle.com/R/entry/analyzing_big_data_using_the1 Here is an example of how we ask ORE to take our IRR_DATA table in the database, split it by the ACCOUNT column, apply a function that calls our SimpleMWRR() calculation, and then combine the results. (If you are following along at home, be sure to have installed our myIRR package on your database server via  “R CMD INSTALL myIRR”). The interesting thing about ore.groupApply is that the calculation is not actually performed in my desktop R environment from which I am running.  What actually happens is that ore.groupApply uses the Oracle database to perform the work.  And the Oracle database is what actually splits the IRR_DATA table by ACCOUNT.  Then the Oracle database takes the data for each account and sends it to an embedded R engine running on the database server to apply our R function.  Then the Oracle database combines all the individual results from the calls to the R function. This is significant because now the embedded R engine only needs to deal with the data for a single account at a time.  Regardless of whether we have 20 accounts or 1 million accounts or more, the R engine that performs the calculation does not care.  Given that normal R has a finite amount of memory to hold data, the ore.groupApply approach overcomes the R memory scalability problem since we only need to fit the data from a single account in R memory (not all of the data for all of the accounts). Additionally, the IRR_DATA does not need to be sent from the database to my desktop R program.  Even though I am invoking ore.groupApply from my desktop R program, because the actual SimpleMWRR calculation is run by the embedded R engine on the database server, the IRR_DATA does not need to leave the database server- this is both a performance benefit because network transmission of large amounts of data take time and a security benefit because it is harder to protect private data once you start shipping around your intranet. Another benefit, which we will discuss in a few paragraphs, is the ability to leverage Oracle database parallelism to run these calculations for dozens of accounts at once. From Walking to Running ore.groupApply is rather nice, but it still has the drawback that I run this from a desktop R instance.  This is not ideal for integrating into typical operational processes like nightly data warehouse refreshes or monthly statement generation.  But, this is not an issue for ORE.  Oracle R Enterprise lets us run this from the database using regular SQL, which is easily integrated into standard operations.  That is extremely exciting and the way we actually did these calculations in the customer proof. As part of Oracle R Enterprise, it provides a SQL equivalent to ore.groupApply which it refers to as “rqGroupEval”.  To use rqGroupEval via SQL, there is a bit of simple setup needed.  Basically, the Oracle Database needs to know the structure of the input table and the grouping column, which we are able to define using the database’s pipeline table function mechanisms. Here is the setup script: At this point, our initial setup of rqGroupEval is done for the IRR_DATA table.  The next step is to define our R function to the database.  We do that via a call to ORE’s rqScriptCreate. Now we can test it.  The SQL you use to run rqGroupEval uses the Oracle database pipeline table function syntax.  The first argument to irr_dataGroupEval is a cursor defining our input.  You can add additional where clauses and subqueries to this cursor as appropriate.  The second argument is any additional inputs to the R function.  The third argument is the text of a dummy select statement.  The dummy select statement is used by the database to identify the columns and datatypes to expect the R function to return.  The fourth argument is the column of the input table to split/group by.  The final argument is the name of the R function as you defined it when you called rqScriptCreate(). The Real-World Results In our real customer proof-of-concept, we had more sophisticated calculation requirements than shown in this simplified blog example.  For instance, we had to perform the rate of return calculations for 5 separate time periods, so the R code was enhanced to do so.  In addition, some accounts needed a time-weighted rate of return to be calculated, so we extended our approach and added an R function to do that.  And finally, there were also a few more real-world data irregularities that we needed to account for, so we added logic to our R functions to deal with those exceptions.  For the full-scale customer test, we loaded the customer data onto a Half-Rack Exadata X2-2 Database Machine.  As our half-rack had 48 physical cores (and 96 threads if you consider hyperthreading), we wanted to take advantage of that CPU horsepower to speed up our calculations.  To do so with ORE, it is as simple as leveraging the Oracle Database Parallel Query features.  Let’s look at the SQL used in the customer proof: Notice that we use a parallel hint on the cursor that is the input to our rqGroupEval function.  That is all we need to do to enable Oracle to use parallel R engines. Here are a few screenshots of what this SQL looked like in the Real-Time SQL Monitor when we ran this during the proof of concept (hint: you might need to right-click on these images to be able to view the images full-screen to see the entire image): From the above, you can notice a few things (numbers 1 thru 5 below correspond with highlighted numbers on the images above.  You may need to right click on the above images and view the images full-screen to see the entire image): The SQL completed in 110 seconds (1.8minutes) We calculated rate of returns for 5 time periods for each of 911k accounts (the number of actual rows returned by the IRRSTAGEGROUPEVAL operation) We accessed 103m rows of detailed cash flow/market value data (the number of actual rows returned by the IRR_STAGE2 operation) We ran with 72 degrees of parallelism spread across 4 database servers Most of our 110seconds was spent in the “External Procedure call” event On average, we performed 8,200 executions of our R function per second (110s/911k accounts) On average, each execution was passed 110 rows of data (103m detail rows/911k accounts) On average, we did 41,000 single time period rate of return calculations per second (each of the 8,200 executions of our R function did rate of return calculations for 5 time periods) On average, we processed over 900,000 rows of database data in R per second (103m detail rows/110s) R + Oracle R Enterprise: Best of R + Best of Oracle Database This blog post series started by describing a real customer problem: how to perform a lot of calculations on a lot of data in a short period of time.  While standard R proved to be a very good fit for writing the necessary calculations, the challenge of working with a lot of data in a short period of time remained. This blog post series showed how Oracle R Enterprise enables R to be used in conjunction with the Oracle Database to overcome the data volume and performance issues (as well as simplifying the operations and security issues).  It also showed that we could calculate 5 time periods of rate of returns for almost a million individual accounts in less than 2 minutes. In a future post, we will take the same R function and show how Oracle R Connector for Hadoop can be used in the Hadoop world.  In that next post, instead of having our data in an Oracle database, our data will live in Hadoop and we will how to use the Oracle R Connector for Hadoop and other Oracle Big Data Connectors to move data between Hadoop, R, and the Oracle Database easily.

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  • Inside the DLR – Invoking methods

    - by Simon Cooper
    So, we’ve looked at how a dynamic call is represented in a compiled assembly, and how the dynamic lookup is performed at runtime. The last piece of the puzzle is how the resolved method gets invoked, and that is the subject of this post. Invoking methods As discussed in my previous posts, doing a full lookup and bind at runtime each and every single time the callsite gets invoked would be far too slow to be usable. The results obtained from the callsite binder must to be cached, along with a series of conditions to determine whether the cached result can be reused. So, firstly, how are the conditions represented? These conditions can be anything; they are determined entirely by the semantics of the language the binder is representing. The binder has to be able to return arbitary code that is then executed to determine whether the conditions apply or not. Fortunately, .NET 4 has a neat way of representing arbitary code that can be easily combined with other code – expression trees. All the callsite binder has to return is an expression (called a ‘restriction’) that evaluates to a boolean, returning true when the restriction passes (indicating the corresponding method invocation can be used) and false when it does’t. If the bind result is also represented in an expression tree, these can be combined easily like so: if ([restriction is true]) { [invoke cached method] } Take my example from my previous post: public class ClassA { public static void TestDynamic() { CallDynamic(new ClassA(), 10); CallDynamic(new ClassA(), "foo"); } public static void CallDynamic(dynamic d, object o) { d.Method(o); } public void Method(int i) {} public void Method(string s) {} } When the Method(int) method is first bound, along with an expression representing the result of the bind lookup, the C# binder will return the restrictions under which that bind can be reused. In this case, it can be reused if the types of the parameters are the same: if (thisArg.GetType() == typeof(ClassA) && arg1.GetType() == typeof(int)) { thisClassA.Method(i); } Caching callsite results So, now, it’s up to the callsite to link these expressions returned from the binder together in such a way that it can determine which one from the many it has cached it should use. This caching logic is all located in the System.Dynamic.UpdateDelegates class. It’ll help if you’ve got this type open in a decompiler to have a look yourself. For each callsite, there are 3 layers of caching involved: The last method invoked on the callsite. All methods that have ever been invoked on the callsite. All methods that have ever been invoked on any callsite of the same type. We’ll cover each of these layers in order Level 1 cache: the last method called on the callsite When a CallSite<T> object is first instantiated, the Target delegate field (containing the delegate that is called when the callsite is invoked) is set to one of the UpdateAndExecute generic methods in UpdateDelegates, corresponding to the number of parameters to the callsite, and the existance of any return value. These methods contain most of the caching, invoke, and binding logic for the callsite. The first time this method is invoked, the UpdateAndExecute method finds there aren’t any entries in the caches to reuse, and invokes the binder to resolve a new method. Once the callsite has the result from the binder, along with any restrictions, it stitches some extra expressions in, and replaces the Target field in the callsite with a compiled expression tree similar to this (in this example I’m assuming there’s no return value): if ([restriction is true]) { [invoke cached method] return; } if (callSite._match) { _match = false; return; } else { UpdateAndExecute(callSite, arg0, arg1, ...); } Woah. What’s going on here? Well, this resulting expression tree is actually the first level of caching. The Target field in the callsite, which contains the delegate to call when the callsite is invoked, is set to the above code compiled from the expression tree into IL, and then into native code by the JIT. This code checks whether the restrictions of the last method that was invoked on the callsite (the ‘primary’ method) match, and if so, executes that method straight away. This means that, the next time the callsite is invoked, the first code that executes is the restriction check, executing as native code! This makes this restriction check on the primary cached delegate very fast. But what if the restrictions don’t match? In that case, the second part of the stitched expression tree is executed. What this section should be doing is calling back into the UpdateAndExecute method again to resolve a new method. But it’s slightly more complicated than that. To understand why, we need to understand the second and third level caches. Level 2 cache: all methods that have ever been invoked on the callsite When a binder has returned the result of a lookup, as well as updating the Target field with a compiled expression tree, stitched together as above, the callsite puts the same compiled expression tree in an internal list of delegates, called the rules list. This list acts as the level 2 cache. Why use the same delegate? Stitching together expression trees is an expensive operation. You don’t want to do it every time the callsite is invoked. Ideally, you would create one expression tree from the binder’s result, compile it, and then use the resulting delegate everywhere in the callsite. But, if the same delegate is used to invoke the callsite in the first place, and in the caches, that means each delegate needs two modes of operation. An ‘invoke’ mode, for when the delegate is set as the value of the Target field, and a ‘match’ mode, used when UpdateAndExecute is searching for a method in the callsite’s cache. Only in the invoke mode would the delegate call back into UpdateAndExecute. In match mode, it would simply return without doing anything. This mode is controlled by the _match field in CallSite<T>. The first time the callsite is invoked, _match is false, and so the Target delegate is called in invoke mode. Then, if the initial restriction check fails, the Target delegate calls back into UpdateAndExecute. This method sets _match to true, then calls all the cached delegates in the rules list in match mode to try and find one that passes its restrictions, and invokes it. However, there needs to be some way for each cached delegate to inform UpdateAndExecute whether it passed its restrictions or not. To do this, as you can see above, it simply re-uses _match, and sets it to false if it did not pass the restrictions. This allows the code within each UpdateAndExecute method to check for cache matches like so: foreach (T cachedDelegate in Rules) { callSite._match = true; cachedDelegate(); // sets _match to false if restrictions do not pass if (callSite._match) { // passed restrictions, and the cached method was invoked // set this delegate as the primary target to invoke next time callSite.Target = cachedDelegate; return; } // no luck, try the next one... } Level 3 cache: all methods that have ever been invoked on any callsite with the same signature The reason for this cache should be clear – if a method has been invoked through a callsite in one place, then it is likely to be invoked on other callsites in the codebase with the same signature. Rather than living in the callsite, the ‘global’ cache for callsite delegates lives in the CallSiteBinder class, in the Cache field. This is a dictionary, typed on the callsite delegate signature, providing a RuleCache<T> instance for each delegate signature. This is accessed in the same way as the level 2 callsite cache, by the UpdateAndExecute methods. When a method is matched in the global cache, it is copied into the callsite and Target cache before being executed. Putting it all together So, how does this all fit together? Like so (I’ve omitted some implementation & performance details): That, in essence, is how the DLR performs its dynamic calls nearly as fast as statically compiled IL code. Extensive use of expression trees, compiled to IL and then into native code. Multiple levels of caching, the first of which executes immediately when the dynamic callsite is invoked. And a clever re-use of compiled expression trees that can be used in completely different contexts without being recompiled. All in all, a very fast and very clever reflection caching mechanism.

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  • Windows XP - Repairing Corrupt System32\Config\System File

    - by SimonTewsi
    My apologies for this long post. I would like to describe the mess I'm in then ask some questions about how to fix it: Starting up my Windows XP SP1 machine I got the following message: Windows could not start because the following file is missing or corrupt: \WINDOWS\SYSTEM32\CONFIG\SYSTEM Tried restarting several times with same results then Googled the problem. Tried the fix described here: http://icrontic.com/articles/repair%5Fwindows%5Fxp (since my CPU does not have XD buffer overflow protection I did not set /NOEXECUTE=OPTIN as OS Load Option). This did not work. I then found another fix for the problem on hardwareanalysis.com: Basically, boot to dos prompt (or recovery console if available) and make backups of the following files:- c:\windows\system32\config\system (to c:\windows\tmp\system.bak) c:\windows\system32\config\software (to c:\windows\tmp\software.bak) c:\windows\system32\config\sam (to c:\windows\tmp\sam.bak) c:\windows\system32\config\security (to c:\windows\tmp\security.bak) c:\windows\system32\config\default (to c:\windows\tmp\default.bak) then delete the above files (not the backups!) then copy the above files in c:\windows\repair to the c:\windows\system32\config directory restart your computer This did work (and I wish I'd done it first, since it was completely reversible, unlike the first method). However, afterwards I found that all the user accounts on the PC were gone. I resurrected them by copying the backed up security file back into the system32\config folder (I may have copied the SAM file from backup as well, I cannot remember clearly now). Now the PC boots up and I can log in. However things are still not right. I tried to alter one of the user accounts and found I could not access the User Accounts in the Control Panel. Microsoft KB 919292 had a fix for the problem. However, the fix failed with a Windows Installer error: The Windows Installer Service could not be accessed. This can occur if you are running Windows in safe mode, or if Windows Installer is not correctly installed. Contact your support personnel for assistance. Windows Installer 3.1 was already installed. I reinstalled it but continued to get the Windows Installer error whenever I tried to run the fix in KB 919292. I have since noticed another three problems: 1) Several applications on the PC no longer run, eg Microsoft Word. Shortcuts no longer seem to do anything and if I run the executables directly (eg for Word by running C:\Program Files\Microsoft Office\Office10\Winword.exe) I get a message similar to: "Microsoft Word has not been installed for the current user. Please run setup to install the application." even though the executable is clearly visible in Windows Explorer (and even though Word actually opens - the error dialog appears after Word has opened. Clicking OK to the error dialog closes Word). 2) One or the other of the two fixes I tried for the original problem caused new user profiles to be created. eg My old user profile under the Documents and Settings folder was Simon. The old one still exists but there is now a new one called Simon.DBQ2515. Obviously the new one is being used because Opera (my browser that still works) no longer sees the bookmarks file under my old profile. 3) Probably as a result of fooling around with the Security file, when I try to boot off the Windows XP CD and run the Recovery Console I am now asked for the administrator password. The only problem is there is no administrator account on the PC. There is one account, LocalAdmin, that has administrative rights but when I entered the password for that account it did not work. It is so long since I originally set up the PC that I cannot remember if the original administrator account ever had a password and, if so, what it was. So, my question is: How can I fix this mess? In particular: 1) Having tried the two fixes linked to above, have I irrepairably damaged the Windows instance, requiring a clean reinstallation of Windows + all applications, or should it be possible to get the machine working correctly again without such drastic measures? 2) Is there any way to get around the administrator password so I can use the Recovery Console again, given that there is no account called "administrator" and the password for the one account with admin privileges does not work (and that, before I started the second fix, I was not asked for an administrator password)? 3) Is there any easy way to fix the problem with the applications that think they are not installed? 4) Is there any easy way to fix the problem of the Windows Installer that does not work, even if reinstalled? Cheers Simon

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  • Does the ulkJSON library have limitations when dealing with base64 in Delphi 7?

    - by Da Gopherboy
    I'm working on a project that is using Delphi 7 to consume RESTful services. We are creating and decoding JSON with the ulkJSON library. Up to this point I've been able to successfully build and send JSON containing a base64 string that exceed 5,160kb. I can verify that the base64 is being received by the services and verify the integrity of the base64 once its there. In addition to sending, I can also receive and successfully decode JSON with a smaller (~ 256KB or less) base64. However I am experiencing some issues on the return trip when larger (~1,024KB+) base64 is involved for some reason. Specifically when attempting to use the following JSON format and function combination: JSON: { "message" : "/9j/4AAQSkZJRgABAQEAYABgAAD...." } Function: function checkResults(JSONFormattedString: String): String; var jsonObject : TlkJSONObject; iteration : Integer; i : Integer; x : Integer; begin jsonObject := TlkJSONobject.Create; // Validate that the JSONFormatted string is not empty. // If it is empty, inform the user/programmer, and exit from this routine. if JSONFormattedString = '' then begin result := 'Error: JSON returned is Null'; jsonObject.Free; exit; end; // Now that we can validate that this string is not empty, we are going to // assume that the string is a JSONFormatted string and attempt to parse it. // // If the string is not a valid JSON object (such as an http status code) // throw an exception informing the user/programmer that an unexpected value // has been passed. And exit from this routine. try jsonObject := TlkJSON.ParseText(JSONFormattedString) as TlkJSONobject; except on e:Exception do begin result := 'Error: No JSON was received from web services'; jsonObject.Free; exit; end; end; // Now that the object has been parsed, lets check the contents. try result := jsonObject.Field['message'].value; jsonObject.Free; exit; except on e:Exception do begin result := 'Error: No Message received from Web Services '+e.message; jsonObject.Free; exit; end; end; end; As mentioned above when using the above function, I am able to get small (256KB and less) base64 strings out of the 'message' field of a JSON object. But for some reason if the received JSON is larger than say 1,024kb the following line seems to just stop in its tracks: jsonObject := TlkJSON.ParseText(JSONFormattedString) as TlkJSONobject; No errors, no results. Following the debugger, I can go into the library, and see that the JSON string being passed is not considered to be JSON despite being in the format listed above. The only difference I can find between calls that work as expected and calls that do not work as expect appears to be the size of base64 being transmitted. Am I missing something completely obvious and should be shot for my code implementation (very possible)? Have I missed some notation regarding the limitations of the ulkJSON library? Any input would be extremely helpful. Thanks in advance stack!

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  • Python: Improving long cumulative sum

    - by Bo102010
    I have a program that operates on a large set of experimental data. The data is stored as a list of objects that are instances of a class with the following attributes: time_point - the time of the sample cluster - the name of the cluster of nodes from which the sample was taken code - the name of the node from which the sample was taken qty1 = the value of the sample for the first quantity qty2 = the value of the sample for the second quantity I need to derive some values from the data set, grouped in three ways - once for the sample as a whole, once for each cluster of nodes, and once for each node. The values I need to derive depend on the (time sorted) cumulative sums of qty1 and qty2: the maximum value of the element-wise sum of the cumulative sums of qty1 and qty2, the time point at which that maximum value occurred, and the values of qty1 and qty2 at that time point. I came up with the following solution: dataset.sort(key=operator.attrgetter('time_point')) # For the whole set sys_qty1 = 0 sys_qty2 = 0 sys_combo = 0 sys_max = 0 # For the cluster grouping cluster_qty1 = defaultdict(int) cluster_qty2 = defaultdict(int) cluster_combo = defaultdict(int) cluster_max = defaultdict(int) cluster_peak = defaultdict(int) # For the node grouping node_qty1 = defaultdict(int) node_qty2 = defaultdict(int) node_combo = defaultdict(int) node_max = defaultdict(int) node_peak = defaultdict(int) for t in dataset: # For the whole system ###################################################### sys_qty1 += t.qty1 sys_qty2 += t.qty2 sys_combo = sys_qty1 + sys_qty2 if sys_combo > sys_max: sys_max = sys_combo # The Peak class is to record the time point and the cumulative quantities system_peak = Peak(time_point=t.time_point, qty1=sys_qty1, qty2=sys_qty2) # For the cluster grouping ################################################## cluster_qty1[t.cluster] += t.qty1 cluster_qty2[t.cluster] += t.qty2 cluster_combo[t.cluster] = cluster_qty1[t.cluster] + cluster_qty2[t.cluster] if cluster_combo[t.cluster] > cluster_max[t.cluster]: cluster_max[t.cluster] = cluster_combo[t.cluster] cluster_peak[t.cluster] = Peak(time_point=t.time_point, qty1=cluster_qty1[t.cluster], qty2=cluster_qty2[t.cluster]) # For the node grouping ##################################################### node_qty1[t.node] += t.qty1 node_qty2[t.node] += t.qty2 node_combo[t.node] = node_qty1[t.node] + node_qty2[t.node] if node_combo[t.node] > node_max[t.node]: node_max[t.node] = node_combo[t.node] node_peak[t.node] = Peak(time_point=t.time_point, qty1=node_qty1[t.node], qty2=node_qty2[t.node]) This produces the correct output, but I'm wondering if it can be made more readable/Pythonic, and/or faster/more scalable. The above is attractive in that it only loops through the (large) dataset once, but unattractive in that I've essentially copied/pasted three copies of the same algorithm. To avoid the copy/paste issues of the above, I tried this also: def find_peaks(level, dataset): def grouping(object, attr_name): if attr_name == 'system': return attr_name else: return object.__dict__[attrname] cuml_qty1 = defaultdict(int) cuml_qty2 = defaultdict(int) cuml_combo = defaultdict(int) level_max = defaultdict(int) level_peak = defaultdict(int) for t in dataset: cuml_qty1[grouping(t, level)] += t.qty1 cuml_qty2[grouping(t, level)] += t.qty2 cuml_combo[grouping(t, level)] = (cuml_qty1[grouping(t, level)] + cuml_qty2[grouping(t, level)]) if cuml_combo[grouping(t, level)] > level_max[grouping(t, level)]: level_max[grouping(t, level)] = cuml_combo[grouping(t, level)] level_peak[grouping(t, level)] = Peak(time_point=t.time_point, qty1=node_qty1[grouping(t, level)], qty2=node_qty2[grouping(t, level)]) return level_peak system_peak = find_peaks('system', dataset) cluster_peak = find_peaks('cluster', dataset) node_peak = find_peaks('node', dataset) For the (non-grouped) system-level calculations, I also came up with this, which is pretty: dataset.sort(key=operator.attrgetter('time_point')) def cuml_sum(seq): rseq = [] t = 0 for i in seq: t += i rseq.append(t) return rseq time_get = operator.attrgetter('time_point') q1_get = operator.attrgetter('qty1') q2_get = operator.attrgetter('qty2') timeline = [time_get(t) for t in dataset] cuml_qty1 = cuml_sum([q1_get(t) for t in dataset]) cuml_qty2 = cuml_sum([q2_get(t) for t in dataset]) cuml_combo = [q1 + q2 for q1, q2 in zip(cuml_qty1, cuml_qty2)] combo_max = max(cuml_combo) time_max = timeline.index(combo_max) q1_at_max = cuml_qty1.index(time_max) q2_at_max = cuml_qty2.index(time_max) However, despite this version's cool use of list comprehensions and zip(), it loops through the dataset three times just for the system-level calculations, and I can't think of a good way to do the cluster-level and node-level calaculations without doing something slow like: timeline = defaultdict(int) cuml_qty1 = defaultdict(int) #...etc. for c in cluster_list: timeline[c] = [time_get(t) for t in dataset if t.cluster == c] cuml_qty1[c] = [q1_get(t) for t in dataset if t.cluster == c] #...etc. Does anyone here at Stack Overflow have suggestions for improvements? The first snippet above runs well for my initial dataset (on the order of a million records), but later datasets will have more records and clusters/nodes, so scalability is a concern. This is my first non-trivial use of Python, and I want to make sure I'm taking proper advantage of the language (this is replacing a very convoluted set of SQL queries, and earlier versions of the Python version were essentially very ineffecient straight transalations of what that did). I don't normally do much programming, so I may be missing something elementary. Many thanks!

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  • Explain the Peak and Flag Algorithm

    - by Isaac Levin
    EDIT Just was pointed that the requirements state peaks cannot be ends of Arrays. So I ran across this site http://codility.com/ Which gives you programming problems and gives you certificates if you can solve them in 2 hours. The very first question is one I have seen before, typically called the Peaks and Flags question. If you are not familiar A non-empty zero-indexed array A consisting of N integers is given. A peak is an array element which is larger than its neighbours. More precisely, it is an index P such that 0 < P < N - 1 and A[P - 1] < A[P] A[P + 1] . For example, the following array A: A[0] = 1 A[1] = 5 A[2] = 3 A[3] = 4 A[4] = 3 A[5] = 4 A[6] = 1 A[7] = 2 A[8] = 3 A[9] = 4 A[10] = 6 A[11] = 2 has exactly four peaks: elements 1, 3, 5 and 10. You are going on a trip to a range of mountains whose relative heights are represented by array A. You have to choose how many flags you should take with you. The goal is to set the maximum number of flags on the peaks, according to certain rules. Flags can only be set on peaks. What's more, if you take K flags, then the distance between any two flags should be greater than or equal to K. The distance between indices P and Q is the absolute value |P - Q|. For example, given the mountain range represented by array A, above, with N = 12, if you take: two flags, you can set them on peaks 1 and 5; three flags, you can set them on peaks 1, 5 and 10; four flags, you can set only three flags, on peaks 1, 5 and 10. You can therefore set a maximum of three flags in this case. Write a function that, given a non-empty zero-indexed array A of N integers, returns the maximum number of flags that can be set on the peaks of the array. For example, given the array above the function should return 3, as explained above. Assume that: N is an integer within the range [1..100,000]; each element of array A is an integer within the range [0..1,000,000,000]. Complexity: expected worst-case time complexity is O(N); expected worst-case space complexity is O(N), beyond input storage (not counting the storage required for input arguments). Elements of input arrays can be modified. So this makes sense, but I failed it using this code public int GetFlags(int[] A) { List<int> peakList = new List<int>(); for (int i = 0; i <= A.Length - 1; i++) { if ((A[i] > A[i + 1] && A[i] > A[i - 1])) { peakList.Add(i); } } List<int> flagList = new List<int>(); int distance = peakList.Count; flagList.Add(peakList[0]); for (int i = 1, j = 0, max = peakList.Count; i < max; i++) { if (Math.Abs(Convert.ToDecimal(peakList[j]) - Convert.ToDecimal(peakList[i])) >= distance) { flagList.Add(peakList[i]); j = i; } } return flagList.Count; } EDIT int[] A = new int[] { 7, 10, 4, 5, 7, 4, 6, 1, 4, 3, 3, 7 }; The correct answer is 3, but my application says 2 This I do not get, since there are 4 peaks (indices 1,4,6,8) and from that, you should be able to place a flag at 2 of the peaks (1 and 6) Am I missing something here? Obviously my assumption is that the beginning or end of an Array can be a peak, is this not the case? If this needs to go in Stack Exchange Programmers, I will move it, but thought dialog here would be helpful. EDIT

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  • Difference between SQL 2005 and SQL 2008 for inserting multiple rows with XML

    - by Sam Dahan
    I am using the following SQL code for inserting multiple rows of data in a table. The data is passed to the stored procedure using an XML variable : INSERT INTO MyTable SELECT SampleTime = T.Item.value('SampleTime[1]', 'datetime'), Volume1 = T.Item.value('Volume1[1]', 'float'), Volume2 = T.Item.value('Volume2[1]', 'float') FROM @xml.nodes('//Root/MyRecord') T(item) I have a whole bunch of unit tests to verify that I am inserting the right information, the right number of records, etc.. when I call the stored procedure. All fine and dandy - that is, until we began to monkey around with the compatibility level of the database. The code above worked beautifully as long as we kept the compatibility level of the DB at 90 (SQL 2005). When we set the compatibility level at 100 (SQL 2008), the unit tests failed, because the stored procedure using the code above times out. The unit tests are dropping the database, re-creating it from scripts, and running the tests on the brand new DB, so it's not - I think - a question of the 'old compatibility level' sticking around. Using the SQL Management studio, I made up a quick test SQL script. Using the same XML chunk, I alter the DB compat level , truncate the table, then use the code above to insert 650 rows. When the level is 90 (SQL 2005), it runs in milliseconds. When the level is 100 (SQL 2008) it sometimes takes over a minute, sometimes runs in milliseconds. I'd appreciate any insight anyone might have into that. EDIT The script takes over a minute to run with my actual data, which has more rows than I show here, is a real table, and has an index. With the following example code, the difference goes between milliseconds and around 5 seconds. --use [master] --ALTER DATABASE MyDB SET compatibility_level =100 use [MyDB] declare @xml xml set @xml = '<?xml version="1.0"?> <Root xmlns:xsi="http://www.w3.org/2001/XMLSchema-instance" xmlns:xsd="http://www.w3.org/2001/XMLSchema"> <Record> <SampleTime>2009-01-24T00:00:00</SampleTime> <Volume1>0</Volume1> <Volume2>0</Volume2> </Record> ..... 653 records, sample time spaced out 4 hours ........ </Root>' DECLARE @myTable TABLE( ID int IDENTITY(1,1) NOT NULL, [SampleTime] [datetime] NOT NULL, [Volume1] [float] NULL, [Volume2] [float] NULL) INSERT INTO @myTable select T.Item.value('SampleTime[1]', 'datetime') as SampleTime, Volume1 = T.Item.value('Volume1[1]', 'float'), Volume2 = T.Item.value('Volume2[1]', 'float') FROM @xml.nodes('//Root/Record') T(item) I uncomment the 2 lines at the top, select them and run just that (the ALTER DATABASE statement), then comment the 2 lines, deselect any text and run the whole thing. When I change from 90 to 100, it runs all the time in 5 seconds (I change the level once, but I run the series several times to see if I have consistent results). When I change from 100 to 90, it runs in milliseconds all the time. Just so you can play with it too. I am using SQL Server 2008 R2 standard edition.

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  • Lock free multiple readers single writer

    - by dummzeuch
    I have got an in memory data structure that is read by multiple threads and written by only one thread. Currently I am using a critical section to make this access threadsafe. Unfortunately this has the effect of blocking readers even though only another reader is accessing it. There are two options to remedy this: use TMultiReadExclusiveWriteSynchronizer do away with any blocking by using a lock free approach For 2. I have got the following so far (any code that doesn't matter has been left out): type TDataManager = class private FAccessCount: integer; FData: TDataClass; public procedure Read(out _Some: integer; out _Data: double); procedure Write(_Some: integer; _Data: double); end; procedure TDataManager.Read(out _Some: integer; out _Data: double); var Data: TDAtaClass; begin InterlockedIncrement(FAccessCount); try // make sure we get both values from the same TDataClass instance Data := FData; // read the actual data _Some := Data.Some; _Data := Data.Data; finally InterlockedDecrement(FAccessCount); end; end; procedure TDataManager.Write(_Some: integer; _Data: double); var NewData: TDataClass; OldData: TDataClass; ReaderCount: integer; begin NewData := TDataClass.Create(_Some, _Data); InterlockedIncrement(FAccessCount); OldData := TDataClass(InterlockedExchange(integer(FData), integer(NewData)); // now FData points to the new instance but there might still be // readers that got the old one before we exchanged it. ReaderCount := InterlockedDecrement(FAccessCount); if ReaderCount = 0 then // no active readers, so we can safely free the old instance FreeAndNil(OldData) else begin /// here is the problem end; end; Unfortunately there is the small problem of getting rid of the OldData instance after it has been replaced. If no other thread is currently within the Read method (ReaderCount=0), it can safely be disposed and that's it. But what can I do if that's not the case? I could just store it until the next call and dispose it there, but Windows scheduling could in theory let a reader thread sleep while it is within the Read method and still has got a reference to OldData. If you see any other problem with the above code, please tell me about it. This is to be run on computers with multiple cores and the above methods are to be called very frequently. In case this matters: I am using Delphi 2007 with the builtin memory manager. I am aware that the memory manager probably enforces some lock anyway when creating a new class but I want to ignore that for the moment. Edit: It may not have been clear from the above: For the full lifetime of the TDataManager object there is only one thread that writes to the data, not several that might compete for write access. So this is a special case of MREW.

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  • QGraphicsView and custom Cursors

    - by Etienne de Martel
    I am trying to make use of a mix of custom cursors and preset cursors for my QGraphicsView. In my implementation we have created a notion of "modes" for the view. Meaning that depending on what "mode" the user is in, different things will happen on the left-click, or left-click drag. Anyway, none of that is the problem, just the context. The problem arises when I try to change the cursor for each mode. For instance, for mode 1 we want to show the regular Arrow cursor, but for mode 2, we want to use a custom pixmap. Seemingly simple we call graphicsview->viewport()->setCursor(Qt::QArrowCursor)  when we are switching to mode 1, and graphicsview->viewport()->setCursor(our custom cursor) for mode 2. Except it doesn't work at all. Firstly, the cursor does not change to the custom cursor. That is the first problem. However, if through another operation the drag mode of the graphics view gets set to ScrollHandDrag, the cursor will switch to the custom cursor once the drag operation is complete. Weird. But the plot thickens... Once we switch to the custom cursor, it can never be changed back to the ArrorCursor no matter how many times we call setCursor(Qt::QArrowCursor). it also doesn't seem to matter whether I call setCursor on the viewport or the graphics view itself. So, just for fun, I added a call to graphicsview->unsetCursor() just before we want to change the cursor, and that at least rectifies the second problem. The cursor changes just fine so long as we do a little HandDragging in between. Better, but certainly not optimal. However it should be noted, that doing the unsetCursor on the viewport doesn't work. it must absolutely be done on the graphicsview - regardless of the fact that we are setting the cursor on the viewport. To completely patch over the problem I have added these two lines after I set the cursor: graphicsview->setDragMode(QGraphicsView::ScrollHandDrag); graphicsview->setDragMode(QGraphicsView::NoDrag); Which works, but ye gads!! So something magical is happening inside these two methods that fixes the problem, but glancing at the code I don't see what. Something to do with the fact that the drag mode is changing the cursor I imagine. Just for completeness, I should also mention that the thing that triggers the mode change, is a QPushButton that has been added to the scene using QGraphicsScene->addWidget(). I don't know if that has anything to do with it, but you never know. I am hoping that either someone could clarify why I need to make these seemingly random calls. I don't think I am doing anything wrong anywhere. Thanks in advance for any help. EDIT: Here is an actual code example with the cursor patches as described above. You can look at and/or download them from the link below. It was a little long to paste here. I included the framework around which the cursors are changed, because I have a funny feeling that that is important somehow. https://gist.github.com/712654 The code where the problem lies is in MyGraphicsView.cpp starting at line 104. This is where the cursor is set in the graphics view. It is exactly as described above. Keep in mind, with the very ugly patches in place the cursors do work - more or less. Without those lines you will see very clearly the problems listed in the post above. Also included in the link, is all the code for a mainWindow that uses the view, etc... the only thing missing are the images I am using. But the images themselves don't matter, any 16x16 pngs will do.

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  • Data munging and data import scripting

    - by morpheous
    I need to write some scripts to carry out some tasks on my server (running Ubuntu server 8.04 TLS). The tasks are to be run periodically, so I will be running the scripts as cron jobs. I have divided the tasks into "group A" and "group B" - because (in my mind at least), they are a bit different. Task Group A import data from a file and possibly reformat it - by reformatting, I mean doing things like santizing the data, possibly normalizing it and or running calculations on 'columns' of the data Import the munged data into a database. For now, I am mostly using mySQL for the vast majority of imports - although some files will be imported into a sqlLite database. Note: The files will be mostly text files, although some of the files are in a binary format (my own proprietary format, written by a C++ application I developed). Task Group B Extract data from the database Perform calculations on the data and either insert or update tables in the database. My coding experience is is primarily as a C/C++ developer, although I have been using PHP as well for the last 2 years or so. I am from a windows background so I am still finding my feet in the linux environment. My question is this - I need to write scripts to perform the tasks I described above. Although I suppose I could write a few C++ applications to be used in the shell scripts, I think it may be better to write them in a scripting language (maybe this is a flawed assumption?). My thinking is that it would be easier to modify thins in a script - no need to rebuild etc for changes to functionality. Additionally, C++ data munging in C++ tends to involve more lines of code than "natural" scripting languages such as Perl, Python etc. Assuming that the majority of people on here agree that scripting is the way to go, herein lies my dilema. Which scripting language to use to perform the tasks above (giving my background). My gut instinct tells me that Perl (shudder) would be the most obvious choice for performing all of the above tasks. BUT (and that is a big BUT). The mere mention of Perl makes my toes curl, as I had a very, very bag experience with it a while back. The syntax seems quite unnatural to me - despite how many times I have tried to learn it - so if possible, I would really like to give it a miss. PHP (which I already know), also am not sure is a good candidate for scripting on the CLI (I have not seen many examples on how to do this etc - so I may be wrong). The last thing I must mention is that IF I have to learn a new language in order to do this, I cannot afford (time constraint) to spend more than a day, in learning the key commands/features required in order to do this (I can always learn the details of the language later, once I have actually deployed the scripts). So, which scripting language would you recommend (PHP, Python, Perl, [insert your favorite here]) - and most importantly WHY?. Or, should I just stick to writing little C++ applications that I call in a shell script?. Lastly, if you have suggested a scripting language, can you please show with a FEW lines (Perl mongers - I'm looking in your direction [nothing to cryptic!] ;) ) how I can use the language you suggested to do what I want to do. Hopefully, the lines you present will convince me that it can be done easily and elegantly in the language you suggested.

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  • Reconciling a new BindingList into a master BindingList using LINQ

    - by Neo
    I have a seemingly simple problem whereby I wish to reconcile two lists so that an 'old' master list is updated by a 'new' list containing updated elements. Elements are denoted by a key property. These are my requirements: All elements in either list that have the same key results in an assignment of that element from the 'new' list over the original element in the 'old' list only if any properties have changed. Any elements in the 'new' list that have keys not in the 'old' list will be added to the 'old' list. Any elements in the 'old' list that have keys not in the 'new' list will be removed from the 'old' list. I found an equivalent problem here - http://stackoverflow.com/questions/161432/ - but it hasn't really been answered properly. So, I came up with an algorithm to iterate through the old and new lists and perform the reconciliation as per the above. Before anyone asks why I'm not just replacing the old list object with the new list object in its entirety, it's for presentation purposes - this is a BindingList bound to a grid on a GUI and I need to prevent refresh artifacts such as blinking, scrollbars moving, etc. So the list object must remain the same, only its updated elements changed. Another thing to note is that the objects in the 'new' list, even if the key is the same and all the properties are the same, are completely different instances to the equivalent objects in the 'old' list, so copying references is not an option. Below is what I've come up with so far - it's a generic extension method for a BindingList. I've put comments in to demonstrate what I'm trying to do. public static class BindingListExtension { public static void Reconcile<T>(this BindingList<T> left, BindingList<T> right, string key) { PropertyInfo piKey = typeof(T).GetProperty(key); // Go through each item in the new list in order to find all updated and new elements foreach (T newObj in right) { // First, find an object in the new list that shares its key with an object in the old list T oldObj = left.First(call => piKey.GetValue(call, null).Equals(piKey.GetValue(newObj, null))); if (oldObj != null) { // An object in each list was found with the same key, so now check to see if any properties have changed and // if any have, then assign the object from the new list over the top of the equivalent element in the old list foreach (PropertyInfo pi in typeof(T).GetProperties()) { if (!pi.GetValue(oldObj, null).Equals(pi.GetValue(newObj, null))) { left[left.IndexOf(oldObj)] = newObj; break; } } } else { // The object in the new list is brand new (has a new key), so add it to the old list left.Add(newObj); } } // Now, go through each item in the old list to find all elements with keys no longer in the new list foreach (T oldObj in left) { // Look for an element in the new list with a key matching an element in the old list if (right.First(call => piKey.GetValue(call, null).Equals(piKey.GetValue(oldObj, null))) == null) { // A matching element cannot be found in the new list, so remove the item from the old list left.Remove(oldObj); } } } } It can be called like this: _oldBindingList.Reconcile(newBindingList, "MyKey") However, I'm looking for perhaps a method of doing the same using LINQ type methods such as GroupJoin<, Join<, Select<, SelectMany<, Intersect<, etc. So far, the problem I've had is that each of these LINQ type methods result in brand new intermediary lists (as a return value) and really, I only want to modify the existing list for all the above reasons. If anyone can help with this, would be most appreciated. If not, no worries, the above method (as it were) will suffice for now. Thanks, Jason

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  • C++ Serial Port Question

    - by Pfeffer
    Problem: I have a hand held device that scans those graphic color barcodes on all packaging. There is a track device that I can use that will slide the device automatically. This track device functions by taking ascii code through a serial port. I need to get this thing to work in FileMaker on a Mac. So no terminal programs, etc... What I've got so far: I bought a Keyspan USB/Serial adapter. Using a program called ZTerm I was successful in sending commands to the device. Example: "C,7^M^J" I was also able to do the same thing in Terminal using this command: screen /dev/tty.KeySerial1 57600 and then type in the same command above(but when I typed in I just hit Control-M and Control-J for the carriage return and line feed) Now I'm writing a plug-in for FileMaker(in C++ of course). I want to get what I did above happen in C++ so when I install that plug-in in FileMaker I can just call one of those functions and have the whole process take place right there. I'm able to connect to the device, but I can't talk to it. It is not responding to anything. I've tried connecting to the device(successfully) using these: FILE *comport; if ((comport = fopen("/dev/tty.KeySerial1", "w")) == NULL){...} and int fd; fd = open("/dev/tty.KeySerial1", O_RDWR | O_NOCTTY | O_NDELAY); This is what I've tried so far in way of talking to the device: fputs ("C,7^M^J",comport); or fprintf(comport,"C,7^M^J"); or char buffer[] = { 'C' , ',' , '7' , '^' , 'M' , '^' , 'J' }; fwrite (buffer , 1 , sizeof(buffer) , comport ); or fwrite('C,7^M^J', 1, 1, comport); Questions: When I connected to the device from Terminal and using ZTerm, I was able to set my baud rate of 57600. I think that may be why it isn't responding here. But I don't know how to do it here.... Does any one know how to do that? I tried this, but it didn't work: comport->BaudRate = 57600; There are a lot of class solutions out there but they all call these include files like termios.h and stdio.h. I don't have these and, for whatever reason, I can't find them to download. I've downloaded a few examples but there are like 20 files in them and they're all calling other files I can't find(like the ones listed above). Do I need to find these and if so where? I just don't know enough about C++ Is there a website where I can download libraries?? Another solution might be to put those terminal commands in C++. Is there a way to do that? So this has been driving me crazy. I'm not a C++ guy, I only know basic programming concepts. Is anyone out there a C++ expert? I ideally I'd like this to just work using functions I already have, like those fwrite, fputs stuff. Thanks!

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  • Templated function with two type parameters fails compile when used with an error-checking macro

    - by SirPentor
    Because someone in our group hates exceptions (let's not discuss that here), we tend to use error-checking macros in our C++ projects. I have encountered an odd compilation failure when using a templated function with two type parameters. There are a few errors (below), but I think the root cause is a warning: warning C4002: too many actual parameters for macro 'BOOL_CHECK_BOOL_RETURN' Probably best explained in code: #include "stdafx.h" template<class A, class B> bool DoubleTemplated(B & value) { return true; } template<class A> bool SingleTemplated(A & value) { return true; } bool NotTemplated(bool & value) { return true; } #define BOOL_CHECK_BOOL_RETURN(expr) \ do \ { \ bool __b = (expr); \ if (!__b) \ { \ return false; \ } \ } while (false) \ bool call() { bool thing = true; // BOOL_CHECK_BOOL_RETURN(DoubleTemplated<int, bool>(thing)); // Above line doesn't compile. BOOL_CHECK_BOOL_RETURN((DoubleTemplated<int, bool>(thing))); // Above line compiles just fine. bool temp = DoubleTemplated<int, bool>(thing); // Above line compiles just fine. BOOL_CHECK_BOOL_RETURN(SingleTemplated<bool>(thing)); BOOL_CHECK_BOOL_RETURN(NotTemplated(thing)); return true; } int _tmain(int argc, _TCHAR* argv[]) { call(); return 0; } Here are the errors, when the offending line is not commented out: 1>------ Build started: Project: test, Configuration: Debug Win32 ------ 1>Compiling... 1>test.cpp 1>c:\junk\temp\test\test\test.cpp(38) : warning C4002: too many actual parameters for macro 'BOOL_CHECK_BOOL_RETURN' 1>c:\junk\temp\test\test\test.cpp(38) : error C2143: syntax error : missing ',' before ')' 1>c:\junk\temp\test\test\test.cpp(38) : error C2143: syntax error : missing ';' before '{' 1>c:\junk\temp\test\test\test.cpp(41) : error C2143: syntax error : missing ';' before '{' 1>c:\junk\temp\test\test\test.cpp(48) : error C2143: syntax error : missing ';' before '{' 1>c:\junk\temp\test\test\test.cpp(49) : error C2143: syntax error : missing ';' before '{' 1>c:\junk\temp\test\test\test.cpp(52) : error C2143: syntax error : missing ';' before '}' 1>c:\junk\temp\test\test\test.cpp(54) : error C2065: 'argv' : undeclared identifier 1>c:\junk\temp\test\test\test.cpp(54) : error C2059: syntax error : ']' 1>c:\junk\temp\test\test\test.cpp(55) : error C2143: syntax error : missing ';' before '{' 1>c:\junk\temp\test\test\test.cpp(58) : error C2143: syntax error : missing ';' before '}' 1>c:\junk\temp\test\test\test.cpp(60) : error C2143: syntax error : missing ';' before '}' 1>c:\junk\temp\test\test\test.cpp(60) : fatal error C1004: unexpected end-of-file found 1>Build log was saved at "file://c:\junk\temp\test\test\Debug\BuildLog.htm" 1>test - 12 error(s), 1 warning(s) ========== Build: 0 succeeded, 1 failed, 0 up-to-date, 0 skipped ========== Any ideas? Thanks!

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  • Scripting with the Sun ZFS Storage 7000 Appliance

    - by Geoff Ongley
    The Sun ZFS Storage 7000 appliance has a user friendly and easy to understand graphical web based interface we call the "BUI" or "Browser User Interface".This interface is very useful for many tasks, but in some cases a script (or workflow) may be more appropriate, such as:Repetitive tasksTasks which work on (or obtain information about) a large number of shares or usersTasks which are triggered by an alert threshold (workflows)Tasks where you want a only very basic input, but a consistent output (workflows)The appliance scripting language is based on ECMAscript 3 (close to javascript). I'm not going to cover ECMAscript 3 in great depth (I'm far from an expert here), but I would like to show you some neat things you can do with the appliance, to get you started based on what I have found from my own playing around.I'm making the assumption you have some sort of programming background, and understand variables, arrays, functions to some extent - but of course if something is not clear, please let me know so I can fix it up or clarify it.Variable Declarations and ArraysVariablesECMAScript is a dynamically and weakly typed language. If you don't know what that means, google is your friend - but at a high level it means we can just declare variables with no specific type and on the fly.For example, I can declare a variable and use it straight away in the middle of my code, for example:projects=list();Which makes projects an array of values that are returned from the list(); function (which is usable in most contexts). With this kind of variable, I can do things like:projects.length (this property on array tells you how many objects are in it, good for for loops etc). Alternatively, I could say:projects=3;and now projects is just a simple number.Should we declare variables like this so loosely? In my opinion, the answer is no - I feel it is a better practice to declare variables you are going to use, before you use them - and given them an initial value. You can do so as follows:var myVariable=0;To demonstrate the ability to just randomly assign and change the type of variables, you can create a simple script at the cli as follows (bold for input):fishy10:> script("." to run)> run("cd /");("." to run)> run ("shares");("." to run)> var projects;("." to run)> projects=list();("." to run)> printf("Number of projects is: %d\n",projects.length);("." to run)> projects=152;("." to run)> printf("Value of the projects variable as an integer is now: %d\n",projects);("." to run)> .Number of projects is: 7Value of the projects variable as an integer is now: 152You can also confirm this behaviour by checking the typeof variable we are dealing with:fishy10:> script("." to run)> run("cd /");("." to run)> run ("shares");("." to run)> var projects;("." to run)> projects=list();("." to run)> printf("var projects is of type %s\n",typeof(projects));("." to run)> projects=152;("." to run)> printf("var projects is of type %s\n",typeof(projects));("." to run)> .var projects is of type objectvar projects is of type numberArraysSo you likely noticed that we have already touched on arrays, as the list(); (in the shares context) stored an array into the 'projects' variable.But what if you want to declare your own array? Easy! This is very similar to Java and other languages, we just instantiate a brand new "Array" object using the keyword new:var myArray = new Array();will create an array called "myArray".A quick example:fishy10:> script("." to run)> testArray = new Array();("." to run)> testArray[0]="This";("." to run)> testArray[1]="is";("." to run)> testArray[2]="just";("." to run)> testArray[3]="a";("." to run)> testArray[4]="test";("." to run)> for (i=0; i < testArray.length; i++)("." to run)> {("." to run)>    printf("Array element %d is %s\n",i,testArray[i]);("." to run)> }("." to run)> .Array element 0 is ThisArray element 1 is isArray element 2 is justArray element 3 is aArray element 4 is testWorking With LoopsFor LoopFor loops are very similar to those you will see in C, java and several other languages. One of the key differences here is, as you were made aware earlier, we can be a bit more sloppy with our variable declarations.The general way you would likely use a for loop is as follows:for (variable; test-case; modifier for variable){}For example, you may wish to declare a variable i as 0; and a MAX_ITERATIONS variable to determine how many times this loop should repeat:var i=0;var MAX_ITERATIONS=10;And then, use this variable to be tested against some case existing (has i reached MAX_ITERATIONS? - if not, increment i using i++);for (i=0; i < MAX_ITERATIONS; i++){ // some work to do}So lets run something like this on the appliance:fishy10:> script("." to run)> var i=0;("." to run)> var MAX_ITERATIONS=10;("." to run)> for (i=0; i < MAX_ITERATIONS; i++)("." to run)> {("." to run)>    printf("The number is %d\n",i);("." to run)> }("." to run)> .The number is 0The number is 1The number is 2The number is 3The number is 4The number is 5The number is 6The number is 7The number is 8The number is 9While LoopWhile loops again are very similar to other languages, we loop "while" a condition is met. For example:fishy10:> script("." to run)> var isTen=false;("." to run)> var counter=0;("." to run)> while(isTen==false)("." to run)> {("." to run)>    if (counter==10) ("." to run)>    { ("." to run)>            isTen=true;   ("." to run)>    } ("." to run)>    printf("Counter is %d\n",counter);("." to run)>    counter++;    ("." to run)> }("." to run)> printf("Loop has ended and Counter is %d\n",counter);("." to run)> .Counter is 0Counter is 1Counter is 2Counter is 3Counter is 4Counter is 5Counter is 6Counter is 7Counter is 8Counter is 9Counter is 10Loop has ended and Counter is 11So what do we notice here? Something has actually gone wrong - counter will technically be 11 once the loop completes... Why is this?Well, if we have a loop like this, where the 'while' condition that will end the loop may be set based on some other condition(s) existing (such as the counter has reached 10) - we must ensure that we  terminate this iteration of the loop when the condition is met - otherwise the rest of the code will be followed which may not be desirable. In other words, like in other languages, we will only ever check the loop condition once we are ready to perform the next iteration, so any other code after we set "isTen" to be true, will still be executed as we can see it was above.We can avoid this by adding a break into our loop once we know we have set the condition - this will stop the rest of the logic being processed in this iteration (and as such, counter will not be incremented). So lets try that again:fishy10:> script("." to run)> var isTen=false;("." to run)> var counter=0;("." to run)> while(isTen==false)("." to run)> {("." to run)>    if (counter==10) ("." to run)>    { ("." to run)>            isTen=true;   ("." to run)>            break;("." to run)>    } ("." to run)>    printf("Counter is %d\n",counter);("." to run)>    counter++;    ("." to run)> }("." to run)> printf("Loop has ended and Counter is %d\n", counter);("." to run)> .Counter is 0Counter is 1Counter is 2Counter is 3Counter is 4Counter is 5Counter is 6Counter is 7Counter is 8Counter is 9Loop has ended and Counter is 10Much better!Methods to Obtain and Manipulate DataGet MethodThe get method allows you to get simple properties from an object, for example a quota from a user. The syntax is fairly simple:var myVariable=get('property');An example of where you may wish to use this, is when you are getting a bunch of information about a user (such as quota information when in a shares context):var users=list();for(k=0; k < users.length; k++){     user=users[k];     run('select ' + user);     var username=get('name');     var usage=get('usage');     var quota=get('quota');...Which you can then use to your advantage - to print or manipulate infomation (you could change a user's information with a set method, based on the information returned from the get method). The set method is explained next.Set MethodThe set method can be used in a simple manner, similar to get. The syntax for set is:set('property','value'); // where value is a string, if it was a number, you don't need quotesFor example, we could set the quota on a share as follows (first observing the initial value):fishy10:shares default/test-geoff> script("." to run)> var currentQuota=get('quota');("." to run)> printf("Current Quota is: %s\n",currentQuota);("." to run)> set('quota','30G');("." to run)> run('commit');("." to run)> currentQuota=get('quota');("." to run)> printf("Current Quota is: %s\n",currentQuota);("." to run)> .Current Quota is: 0Current Quota is: 32212254720This shows us using both the get and set methods as can be used in scripts, of course when only setting an individual share, the above is overkill - it would be much easier to set it manually at the cli using 'set quota=3G' and then 'commit'.List MethodThe list method can be very powerful, especially in more complex scripts which iterate over large amounts of data and manipulate it if so desired. The general way you will use list is as follows:var myVar=list();Which will make "myVar" an array, containing all the objects in the relevant context (this could be a list of users, shares, projects, etc). You can then gather or manipulate data very easily.We could list all the shares and mountpoints in a given project for example:fishy10:shares another-project> script("." to run)> var shares=list();("." to run)> for (i=0; i < shares.length; i++)("." to run)> {("." to run)>    run('select ' + shares[i]);("." to run)>    var mountpoint=get('mountpoint');("." to run)>    printf("Share %s discovered, has mountpoint %s\n",shares[i],mountpoint);("." to run)>    run('done');("." to run)> }("." to run)> .Share and-another discovered, has mountpoint /export/another-project/and-anotherShare another-share discovered, has mountpoint /export/another-project/another-shareShare bob discovered, has mountpoint /export/another-projectShare more-shares-for-all discovered, has mountpoint /export/another-project/more-shares-for-allShare yep discovered, has mountpoint /export/another-project/yepWriting More Complex and Re-Usable CodeFunctionsThe best way to be able to write more complex code is to use functions to split up repeatable or reusable sections of your code. This also makes your more complex code easier to read and understand for other programmers.We write functions as follows:function functionName(variable1,variable2,...,variableN){}For example, we could have a function that takes a project name as input, and lists shares for that project (assuming we're already in the 'project' context - context is important!):function getShares(proj){        run('select ' + proj);        shares=list();        printf("Project: %s\n", proj);        for(j=0; j < shares.length; j++)        {                printf("Discovered share: %s\n",shares[i]);        }        run('done'); // exit selected project}Commenting your CodeLike any other language, a large part of making it readable and understandable is to comment it. You can use the same comment style as in C and Java amongst other languages.In other words, sngle line comments use://at the beginning of the comment.Multi line comments use:/*at the beginning, and:*/ at the end.For example, here we will use both:fishy10:> script("." to run)> // This is a test comment("." to run)> printf("doing some work...\n");("." to run)> /* This is a multi-line("." to run)> comment which I will span across("." to run)> three lines in total */("." to run)> printf("doing some more work...\n");("." to run)> .doing some work...doing some more work...Your comments do not have to be on their own, they can begin (particularly with single line comments this is handy) at the end of a statement, for examplevar projects=list(); // The variable projects is an array containing all projects on the system.Try and Catch StatementsYou may be used to using try and catch statements in other languages, and they can (and should) be utilised in your code to catch expected or unexpected error conditions, that you do NOT wish to stop your code from executing (if you do not catch these errors, your script will exit!):try{  // do some work}catch(err) // Catch any error that could occur{ // do something here under the error condition}For example, you may wish to only execute some code if a context can be reached. If you can't perform certain actions under certain circumstances, that may be perfectly acceptable.For example if you want to test a condition that only makes sense when looking at a SMB/NFS share, but does not make sense when you hit an iscsi or FC LUN, you don't want to stop all processing of other shares you may not have covered yet.For example we may wish to obtain quota information on all shares for all users on a share (but this makes no sense for a LUN):function getShareQuota(shar) // Get quota for each user of this share{        run('select ' + shar);        printf("  SHARE: %s\n", shar);        try        {                run('users');                printf("    %20s        %11s    %11s    %3s\n","Username","Usage(G)","Quota(G)","Quota(%)");                printf("    %20s        %11s    %11s    %4s\n","--------","--------","--------","----");                                users=list();                for(k=0; k < users.length; k++)                {                        user=users[k];                        getUserQuota(user);                }                run('done'); // exit user context        }        catch(err)        {                printf("    SKIPPING %s - This is NOT a NFS or CIFs share, not looking for users\n", shar);        }        run('done'); // done with this share}Running Scripts Remotely over SSHAs you have likely noticed, writing and running scripts for all but the simplest jobs directly on the appliance is not going to be a lot of fun.There's a couple of choices on what you can do here:Create scripts on a remote system and run them over sshCreate scripts, wrapping them in workflow code, so they are stored on the appliance and can be triggered under certain circumstances (like a threshold being reached)We'll cover the first one here, and then cover workflows later on (as these are for the most part just scripts with some wrapper information around them).Creating a SSH Public/Private SSH Key PairLog on to your handy Solaris box (You wouldn't be using any other OS, right? :P) and use ssh-keygen to create a pair of ssh keys. I'm storing this separate to my normal key:[geoff@lightning ~] ssh-keygen -t rsa -b 1024Generating public/private rsa key pair.Enter file in which to save the key (/export/home/geoff/.ssh/id_rsa): /export/home/geoff/.ssh/nas_key_rsaEnter passphrase (empty for no passphrase): Enter same passphrase again: Your identification has been saved in /export/home/geoff/.ssh/nas_key_rsa.Your public key has been saved in /export/home/geoff/.ssh/nas_key_rsa.pub.The key fingerprint is:7f:3d:53:f0:2a:5e:8b:2d:94:2a:55:77:66:5c:9b:14 geoff@lightningInstalling the Public Key on the ApplianceOn your Solaris host, observe the public key:[geoff@lightning ~] cat .ssh/nas_key_rsa.pub ssh-rsa AAAAB3NzaC1yc2EAAAABIwAAAIEAvYfK3RIaAYmMHBOvyhKM41NaSmcgUMC3igPN5gUKJQvSnYmjuWG6CBr1CkF5UcDji7v19jG3qAD5lAMFn+L0CxgRr8TNaAU+hA4/tpAGkjm+dKYSyJgEdMIURweyyfUFXoerweR8AWW5xlovGKEWZTAfvJX9Zqvh8oMQ5UJLUUc= geoff@lightningNow, copy and paste everything after "ssh-rsa" and before "user@hostname" - in this case, geoff@lightning. That is, this bit:AAAAB3NzaC1yc2EAAAABIwAAAIEAvYfK3RIaAYmMHBOvyhKM41NaSmcgUMC3igPN5gUKJQvSnYmjuWG6CBr1CkF5UcDji7v19jG3qAD5lAMFn+L0CxgRr8TNaAU+hA4/tpAGkjm+dKYSyJgEdMIURweyyfUFXoerweR8AWW5xlovGKEWZTAfvJX9Zqvh8oMQ5UJLUUc=Logon to your appliance and get into the preferences -> keys area for this user (root):[geoff@lightning ~] ssh [email protected]: Last login: Mon Dec  6 17:13:28 2010 from 192.168.0.2fishy10:> configuration usersfishy10:configuration users> select rootfishy10:configuration users root> preferences fishy10:configuration users root preferences> keysOR do it all in one hit:fishy10:> configuration users select root preferences keysNow, we create a new public key that will be accepted for this user and set the type to RSA:fishy10:configuration users root preferences keys> createfishy10:configuration users root preferences key (uncommitted)> set type=RSASet the key itself using the string copied previously (between ssh-rsa and user@host), and set the key ensuring you put double quotes around it (eg. set key="<key>"):fishy10:configuration users root preferences key (uncommitted)> set key="AAAAB3NzaC1yc2EAAAABIwAAAIEAvYfK3RIaAYmMHBOvyhKM41NaSmcgUMC3igPN5gUKJQvSnYmjuWG6CBr1CkF5UcDji7v19jG3qAD5lAMFn+L0CxgRr8TNaAU+hA4/tpAGkjm+dKYSyJgEdMIURweyyfUFXoerweR8AWW5xlovGKEWZTAfvJX9Zqvh8oMQ5UJLUUc="Now set the comment for this key (do not use spaces):fishy10:configuration users root preferences key (uncommitted)> set comment="LightningRSAKey" Commit the new key:fishy10:configuration users root preferences key (uncommitted)> commitVerify the key is there:fishy10:configuration users root preferences keys> lsKeys:NAME     MODIFIED              TYPE   COMMENT                                  key-000  2010-10-25 20:56:42   RSA    cycloneRSAKey                           key-001  2010-12-6 17:44:53    RSA    LightningRSAKey                         As you can see, we now have my new key, and a previous key I have created on this appliance.Running your Script over SSH from a Remote SystemHere I have created a basic test script, and saved it as test.ecma3:[geoff@lightning ~] cat test.ecma3 script// This is a test script, By Geoff Ongley 2010.printf("Testing script remotely over ssh\n");.Now, we can run this script remotely with our keyless login:[geoff@lightning ~] ssh -i .ssh/nas_key_rsa root@fishy10 < test.ecma3Pseudo-terminal will not be allocated because stdin is not a terminal.Testing script remotely over sshPutting it Together - An Example Completed Quota Gathering ScriptSo now we have a lot of the basics to creating a script, let us do something useful, like, find out how much every user is using, on every share on the system (you will recognise some of the code from my previous examples): script/************************************** Quick and Dirty Quota Check script ** Written By Geoff Ongley            ** 25 October 2010                    **************************************/function getUserQuota(usr){        run('select ' + usr);        var username=get('name');        var usage=get('usage');        var quota=get('quota');        var usage_g=usage / 1073741824; // convert bytes to gigabytes        var quota_g=quota / 1073741824; // as above        var quota_percent=0        if (quota > 0)        {                quota_percent=(usage / quota)*(100/1);        }        printf("    %20s        %8.2f           %8.2f           %d%%\n",username,usage_g,quota_g,quota_percent);        run('done'); // done with this selected user}function getShareQuota(shar){        //printf("DEBUG: selecting share %s\n", shar);        run('select ' + shar);        printf("  SHARE: %s\n", shar);        try        {                run('users');                printf("    %20s        %11s    %11s    %3s\n","Username","Usage(G)","Quota(G)","Quota(%)");                printf("    %20s        %11s    %11s    %4s\n","--------","--------","--------","--------");                                users=list();                for(k=0; k < users.length; k++)                {                        user=users[k];                        getUserQuota(user);                }                run('done'); // exit user context        }        catch(err)        {                printf("    SKIPPING %s - This is NOT a NFS or CIFs share, not looking for users\n", shar);        }        run('done'); // done with this share}function getShares(proj){        //printf("DEBUG: selecting project %s\n",proj);        run('select ' + proj);        shares=list();        printf("Project: %s\n", proj);        for(j=0; j < shares.length; j++)        {                share=shares[j];                getShareQuota(share);        }        run('done'); // exit selected project}function getProjects(){        run('cd /');        run('shares');        projects=list();                for (i=0; i < projects.length; i++)        {                var project=projects[i];                getShares(project);        }        run('done'); // exit context for all projects}getProjects();.Which can be run as follows, and will print information like this:[geoff@lightning ~/FISHWORKS_SCRIPTS] ssh -i ~/.ssh/nas_key_rsa root@fishy10 < get_quota_utilisation.ecma3Pseudo-terminal will not be allocated because stdin is not a terminal.Project: another-project  SHARE: and-another                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                  nobody            0.00            0.00        0%                 geoffro            0.05            0.00        0%                   Billy            0.10            0.00        0%                    root            0.00            0.00        0%            testing-user            0.05            0.00        0%  SHARE: another-share                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                    root            0.00            0.00        0%                  nobody            0.00            0.00        0%                 geoffro            0.05            0.49        9%            testing-user            0.05            0.02        249%                   Billy            0.10            0.29        33%  SHARE: bob                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                  nobody            0.00            0.00        0%                    root            0.00            0.00        0%  SHARE: more-shares-for-all                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                   Billy            0.10            0.00        0%            testing-user            0.05            0.00        0%                  nobody            0.00            0.00        0%                    root            0.00            0.00        0%                 geoffro            0.05            0.00        0%  SHARE: yep                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                    root            0.00            0.00        0%                  nobody            0.00            0.00        0%                   Billy            0.10            0.01        999%            testing-user            0.05            0.49        9%                 geoffro            0.05            0.00        0%Project: default  SHARE: Test-LUN    SKIPPING Test-LUN - This is NOT a NFS or CIFs share, not looking for users  SHARE: test-geoff                Username           Usage(G)       Quota(G)    Quota(%)                --------           --------       --------    --------                 geoffro            0.05            0.00        0%                    root            3.18           10.00        31%                    uucp            0.00            0.00        0%                  nobody            0.59            0.49        119%^CKilled by signal 2.Creating a WorkflowWorkflows are scripts that we store on the appliance, and can have the script execute either on request (even from the BUI), or on an event such as a threshold being met.Workflow BasicsA workflow allows you to create a simple process that can be executed either via the BUI interface interactively, or by an alert being raised (for some threshold being reached, for example).The basics parameters you will have to set for your "workflow object" (notice you're creating a variable, that embodies ECMAScript) are as follows (parameters is optional):name: A name for this workflowdescription: A Description for the workflowparameters: A set of input parameters (useful when you need user input to execute the workflow)execute: The code, the script itself to execute, which will be function (parameters)With parameters, you can specify things like this (slightly modified sample taken from the System Administration Guide):          ...parameters:        variableParam1:         {                             label: 'Name of Share',                             type: 'String'                  },                  variableParam2                  {                             label: 'Share Size',                             type: 'size'                  },execute: ....};  Note the commas separating the sections of name, parameters, execute, and so on. This is important!Also - there is plenty of properties you can set on the parameters for your workflow, these are described in the Sun ZFS Storage System Administration Guide.Creating a Basic Workflow from a Basic ScriptTo make a basic script into a basic workflow, you need to wrap the following around your script to create a 'workflow' object:var workflow = {name: 'Get User Quotas',description: 'Displays Quota Utilisation for each user on each share',execute: function() {// (basic script goes here, minus the "script" at the beginning, and "." at the end)}};However, it appears (at least in my experience to date) that the workflow object may only be happy with one function in the execute parameter - either that or I'm doing something wrong. As far as I can tell, after execute: you should only have a basic one function context like so:execute: function(){}To deal with this, and to give an example similar to our script earlier, I have created another simple quota check, to show the same basic functionality, but in a workflow format:var workflow = {name: 'Get User Quotas',description: 'Displays Quota Utilisation for each user on each share',execute: function () {        run('cd /');        run('shares');        projects=list();                for (i=0; i < projects.length; i++)        {                run('select ' + projects[i]);                shares=list('filesystem');                printf("Project: %s\n", projects[i]);                for(j=0; j < shares.length; j++)                {                        run('select ' +shares[j]);                        try                        {                                run('users');                                printf("  SHARE: %s\n", shares[j]);                                printf("    %20s        %11s    %11s    %3s\n","Username","Usage(G)","Quota(G)","Quota(%)");                                printf("    %20s        %11s    %11s    %4s\n","--------","--------","--------","-------");                                users=list();                                for(k=0; k < users.length; k++)                                {                                        run('select ' + users[k]);                                        username=get('name');                                        usage=get('usage');                                        quota=get('quota');                                        usage_g=usage / 1073741824; // convert bytes to gigabytes                                        quota_g=quota / 1073741824; // as above                                        quota_percent=0                                        if (quota > 0)                                        {                                                quota_percent=(usage / quota)*(100/1);                                        }                                        printf("    %20s        %8.2f   %8.2f   %d%%\n",username,usage_g,quota_g,quota_percent);                                        run('done');                                }                                run('done'); // exit user context                        }                        catch(err)                        {                        //      printf("    %s is a LUN, Not looking for users\n", shares[j]);                        }                        run('done'); // exit selected share context                }                run('done'); // exit project context        }        }};SummaryThe Sun ZFS Storage 7000 Appliance offers lots of different and interesting features to Sun/Oracle customers, including the world renowned Analytics. Hopefully the above will help you to think of new creative things you could be doing by taking advantage of one of the other neat features, the internal scripting engine!Some references are below to help you continue learning more, I'll update this post as I do the same! Enjoy...More information on ECMAScript 3A complete reference to ECMAScript 3 which will help you learn more of the details you may be interested in, can be found here:http://www.ecma-international.org/publications/files/ECMA-ST-ARCH/ECMA-262,%203rd%20edition,%20December%201999.pdfMore Information on Administering the Sun ZFS Storage 7000The Sun ZFS Storage 7000 System Administration guide can be a useful reference point, and can be found here:http://wikis.sun.com/download/attachments/186238602/2010_Q3_2_ADMIN.pdf

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  • Ninject.Web, OnePerRequestModule, and IIS7 Integrated Pipeline

    - by Ted
    Using Ninject.Web with ASP.NET WebForms project. Works without issues using classic pipeline, but when it's under integrated pipeline, a null reference exception occurs on every request (which I've narrowed down to the use of the OnePerRequestModule): [NullReferenceException: Object reference not set to an instance of an object.] System.Web.PipelineStepManager.ResumeSteps(Exception error) +1216 System.Web.HttpApplication.BeginProcessRequestNotification(HttpContext context, AsyncCallback cb) +113 System.Web.HttpRuntime.ProcessRequestNotificationPrivate(IIS7WorkerRequest wr, HttpContext context) +616 The above always occurs unless I remove the OnePerRequestModule initializization. occurs consistently on a very basic test app I put together. On a standard app where I actually want to implement it, I can solve the issue by initializing the OnePerRequestModule like so: protected override IKernel CreateKernel() { // This will always blow up. //var module = new OnePerRequestModule(); //module.Init(this); IKernel kernel = new StandardKernel(new MyModule()); // This works on larger app, but on basic app, it makes no difference under integrated pipeline as the above exception is always thrown. var module = new OnePerRequestModule(); module.Init(this); return kernel; } Before I start spelunking further, is anybody out there using Ninject.Web extension successfully under the integrated pipeline in IIS7 AND using the OnePerRequestModule? There are certain restrictions for modules under the integrated pipeline that weren't there in previous IIS versions/classic pipeline. Quickly thrown together sample project at http://www.filedropper.com/test_59 And in case it's not obvious with Ninject.Web: it's an ASP.NET WebForms project.

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  • [WPF] Custom TabItem in TabControl

    - by Simon
    I've created CustomTabItem which inherits from TabItem and i'd like to use it while binding ObservableCollection in TabControl <TabControl ItemsSource="{Binding MyObservableCollection}"/> It should like this in XAML, but i do not know how change default type of the output item created by TabControl while binding. I've tried to create converter, but it has to do something like this inside convertin method: List<CustomTabItem> resultList = new List<CustomTabItem>(); And iterate through my input ObservableCollection, create my CustomTab based on item from collection and add it to resultList... I'd like to avoid it, bacause while creating CustomTabItem i'm creating complex View and it takes a while, so i don't want to create it always when something change in binded collection. My class extends typical TabItem and i'd like to use this class in TabControl instead of TabItem. <TabControl.ItemContainerStyle> <Style TargetType="{x:Type local:CustomTabItem}"> <Setter Property="MyProperty" Value="{Binding xxx}"/> </Style> </TabControl.ItemContainerStyle> Code above generates error that Style cannot be applied to TabItem. My main purpose is to use in XAML my own CustomTabItem and bind properties... Just like above... I've also tried to use <TabControl.ItemTemplate/> <TabControl.ContentTemaplte/> But they are just styles for TabItem, so i'll still be missing my properties wchich i added in my custom class.

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  • NServiceBus and NHibernate - Message Handler and Transactions

    - by mattcodes
    From my understanding NServiceBus executes the Handle method of an IMessageHandler within a transaction, if an exception propagates out of this method, then NServiceBus will ensure the message is put back on the message queue (up X amount of times before error queue) etc.. so we have an atomic operation so to speak. Now when if I inside my NServiceBus Message Handle method I do something like this using(var trans = session.BeginTransaction()) { person.Age = 10; session.Update<Person>(person); trans.Commit() } using(var trans2 = session.BeginTransaction()) { person.Age = 20; session.Update<Person>(person); // throw new ApplicationException("Oh no"); trans2.Commit() } What is the effect of this on the transaction scope? Is trans1 now counted as a nested transaction in terms of its relationship with the Nservicebus transaction even though we have done nothing to marry them up? (if not how would one link onto the transaction of NServiceBus? Looking at the second block (trans2), if I uncomment the throw statement, will the NServiceBus transaction then rollback trans1 as well? In basic scenarios, say I dump the above into a console app, then trans1 is independent, commit, flushed and won't rollback. I'm trying to clarify what happens now we sit in someone else's transaction like NServiceBus? The above is just example code, im wouldnt be working directly with session, more like through a uow pattern.

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  • -[UIImage drawInRect:] / CGContextDrawImage() not releasing memory?

    - by sohocoke
    I wanted to easily blend a UIImage on top of another background image, so wrote a category method for UIImage, adapted from http://stackoverflow.com/questions/1309757/blend-two-uiimages : - (UIImage *) blendedImageOn:(UIImage *) backgroundImage { NSAutoreleasePool* pool = [[NSAutoreleasePool alloc] init]; UIGraphicsBeginImageContext(backgroundImage.size); CGRect rect = CGRectMake(0, 0, backgroundImage.size.width, backgroundImage.size.height); [backgroundImage drawInRect:rect]; [self drawInRect:rect]; UIImage* blendedImage = UIGraphicsGetImageFromCurrentImageContext(); UIGraphicsEndImageContext(); [pool release]; return blendedImage; } Unfortunately my app that uses the above method to load around 20 images and blend them with background and gloss images (so probably around 40 calls), is being jettisoned on the device. An Instruments session revealed that calls to malloc stemming from the calls to drawInRect: are responsible for the bulk of the memory usage. I tried replacing the drawInRect: messages with equivalent function calls to the function CGContextDrawImage but it didn't help. The AutoReleasePool was added after I found the memory usage problem; it also didn't make a difference. I'm thinking this is probably because I'm not using graphics contexts appropriately. Would calling the above method in a loop be a bad idea because of the number of contexts I create? Or did I simply miss something?

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