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• Monitor and Control Memory Usage in Google Chrome

  - by Asian Angel

  Do you want to know just how much memory Google Chrome and any installed extensions are using at a given moment? With just a few clicks you can see just what is going on under the hood of your browser. How Much Memory are the Extensions Using? Here is our test browser with a new tab and the Extensions Page open, five enabled extensions, and one disabled at the moment. You can access Chrome’s Task Manager using the Page Menu, going to Developer, and selecting Task manager… Or by right clicking on the Tab Bar and selecting Task manager. There is also a keyboard shortcut (Shift + Esc) available for the “keyboard ninjas”. Sitting idle as shown above here are the stats for our test browser. All of the extensions are sitting there eating memory even though some of them are not available/active for use on our new tab and Extensions Page. Not so good… If the default layout is not to your liking then you can easily modify the information that is available by right clicking and adding/removing extra columns as desired. For our example we added Shared Memory & Private Memory. Using the about:memory Page to View Memory Usage Want even more detail? Type about:memory into the Address Bar and press Enter. Note: You can also access this page by clicking on the Stats for nerds Link in the lower left corner of the Task Manager Window. Focusing on the four distinct areas you can see the exact version of Chrome that is currently installed on your system… View the Memory & Virtual Memory statistics for Chrome… Note: If you have other browsers running at the same time you can view statistics for them here too. See a list of the Processes currently running… And the Memory & Virtual Memory statistics for those processes. The Difference with the Extensions Disabled Just for fun we decided to disable all of the extension in our test browser… The Task Manager Window is looking rather empty now but the memory consumption has definitely seen an improvement. Comparing Memory Usage for Two Extensions with Similar Functions For our next step we decided to compare the memory usage for two extensions with similar functionality. This can be helpful if you are wanting to keep memory consumption trimmed down as much as possible when deciding between similar extensions. First up was Speed Dial”(see our review here). The stats for Speed Dial…quite a change from what was shown above (~3,000 – 6,000 K). Next up was Incredible StartPage (see our review here). Surprisingly both were nearly identical in the amount of memory being used. Purging Memory Perhaps you like the idea of being able to “purge” some of that excess memory consumption. With a simple command switch modification to Chrome’s shortcut(s) you can add a Purge Memory Button to the Task Manager Window as shown below.  Notice the amount of memory being consumed at the moment… Note: The tutorial for adding the command switch can be found here. One quick click and there is a noticeable drop in memory consumption. Conclusion We hope that our examples here will prove useful to you in managing the memory consumption in your own Google Chrome installation. If you have a computer with limited resources every little bit definitely helps out. Similar Articles Productive Geek Tips Stupid Geek Tricks: Compare Your Browser’s Memory Usage with Google ChromeMonitor CPU, Memory, and Disk IO In Windows 7 with Taskbar MetersFix for Firefox memory leak on WindowsHow to Purge Memory in Google ChromeHow to Make Google Chrome Your Default Browser TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Acronis Online Backup DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows iFixit Offers Gadget Repair Manuals Online Vista style sidebar for Windows 7 Create Nice Charts With These Web Based Tools Track Daily Goals With 42Goals Video Toolbox is a Superb Online Video Editor Fun with 47 charts and graphs

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• Measuring Code Quality

  - by DotNetBlues

  Several months back, I was tasked with measuring the quality of code in my organization. Foolishly, I said, "No problem." I figured that Visual Studio has a built-in code metrics tool (Analyze -> Calculate Code Metrics) and that would be a fine place to start with. I was right, but also very wrong. The Visual Studio calculates five primary metrics: Maintainability Index, Cyclomatic Complexity, Depth of Inheritance, Class Coupling, and Lines of Code. The first two are figured at the method level, the second at (primarily) the class level, and the last is a simple count. The first question any reasonable person should ask is "Which one do I look at first?" The first question any manager is going to ask is, "What one number tells me about the whole application?" My answer to both, in a way, was "Maintainability Index." Why? Because each of the other numbers represent one element of quality while MI is a composite number that includes Cyclomatic Complexity. I'd be lying if I said no consideration was given to the fact that it was abstract enough that it's harder for some surly developer (I've been known to resemble that remark) to start arguing why a high coupling or inheritance is no big deal or how complex requirements are to blame for complex code. I should also note that I don't think there is one magic bullet metric that will tell you objectively how good a code base is. There are a ton of different metrics out there, and each one was created for a specific purpose in mind and has a pet theory behind it. When you've got a group of developers who aren't accustomed to measuring code quality, picking a 0-100 scale, non-controversial metric that can be easily generated by tools you already own really isn't a bad place to start. That sort of answers the question a developer would ask, but what about the management question; how do you dashboard this stuff when Visual Studio doesn't roll up the numbers to the solution level? Since VS does roll up the MI to the project level, I thought I could just figure out what sort of weighting Microsoft used to roll method scores up to the class level and then to the namespace and project levels. I was a bit surprised by the answer: there is no weighting. That means that a class with one 1300 line method (which will score a 0 MI) and one empty constructor (which will score a 100 MI) will have an overall MI of a respectable 50. Throw in a couple of DTOs that are nothing more than getters and setters (which tend to score 95 or better) and the project ends up looking really, really healthy. The next poor bastard who has to work on the application is probably not going to be singing the praises of its maintainability, though. For the record, that 1300 line method isn't a hypothetical, either. So, what does one do with that? Well, I decided to weight the average by the Lines of Code per method. For our above example, the formula for the class's MI becomes ((1300 * 0) + (1 * 100))/1301 = .077, rounded to 0. Sounds about right. Continue the pattern for namespace, project, solution, and even multi-solution application MI scores. This can be done relatively easily by using the "export to Excel" button and running a quick formula against the data. On the short list of follow-up questions would be, "How do I improve my application's score?" That's an answer for another time, though.

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• The Latest News About SAP

  - by jmorourke

  Like many professionals, I get a lot of my news from Google e-mail alerts that I’ve set up to keep track of key industry trends and competitive news.  In the past few weeks, I’ve been getting a number of news alerts about SAP.  Below are a few recent examples: Warm weather cuts short US maple sugaring season – by Toby Talbot, AP MILWAUKEE – Temperatures in Wisconsin had already hit the high 60s when Gretchen Grape and her family began tapping their 850 maple trees. They had waited for the state's ceremonial tapping to kick off the maple sugaring season. It was moved up five days, but that didn't make much difference. For Grape, the typically month-long season ended nine days later. The SAP had stopped flowing in a record-setting heat wave, and the 5-quart collection bags that in a good year fill in a day were still half-empty. Instead of their usual 300 gallons of syrup, her family had about 40. Maple syrup producers across the North have had their season cut short by unusually warm weather. While those with expensive, modern vacuum systems say they've been able to suck a decent amount of sap from their trees, producers like Grape, who still rely on traditional taps and buckets, have seen their year ruined. "It's frustrating," said the 69-year-old retiree from Holcombe, Wis. "You put in the same amount of work, equipment, investment, and then all of a sudden, boom, you have no SAP." Home & Garden: Too-Early Spring Means Sugaring Woes  - by Georgeanne Davis for The Free Press Over this past weekend, forsythia and daffodils were blooming in the southern parts of the state as temperatures climbed to 85 degrees, and trees began budding out, putting an end to this year's maple syrup production even as the state celebrated Maine Maple Sunday. Maple sugaring needs cold nights and warm days to induce SAP flows. Once the trees begin budding, SAP can still flow, but the SAP is bitter and has an off taste. Many farmers and dairymen count on sugaring for extra income, so the abbreviated season is a real financial loss for them, akin to the shortened shrimping season's effect on Maine lobstermen. SAP season comes to a sugary Sunday finale – Kennebec Journal, March 26th, 2012 Rebecca Manthey stood out in the rain at the entrance of Old Fort Western keeping watch over a cast iron kettle of boiling SAP hooked to a tripod over a wood fire.  Manthey and the rest of the Old Fort Western staff -- decked out in 18th-century attire -- joined sugar houses across the state in observance of Maine Maple Sunday. The annual event is sponsored by the Department of Agriculture and the Maine Maple Producers Association.  She said the rain hadn't kept people from coming to enjoy all the events at the fort surrounding the production of Maple syrup.  "In the 18th century, you would be boiling SAP in the woods, so I would be in the woods," Manthey explained to the families who circled around her. "People spent weeks and weeks in the woods. You don't want to cook it to fast or it would burn. When it looks like the right consistency then you send it (into the kitchen) to be made into sugar." Manthey said she enjoyed portraying an 18th-century woman, even in the rain, which didn't seem to bother visitors either. There was a steady stream of families touring the fort and enjoying the maple syrup demonstrations. I hope you enjoy these updates on SAP – Happy April Fool’s Day!

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• BizTalk 2009 - Pipeline Component Wizard

  - by Stuart Brierley

  Recently I decided to try out the BizTalk Server Pipeline Component Wizard when creating a new pipeline component for BizTalk 2009. There are different versions of the wizard available, so be sure to download the appropriate version for the BizTalk environment that you are working with. Following the download and expansion of the zip file, you should be left with a Visual Studio solution.  Open this solution and build the project. Following this installation is straight foward - locate and run the built setup.exe file in the PipelineComponentWizard Setup project and click through the small number of installation screens. Once you have completed installation you will be ready to use the wizard in Visual Studio to create your BizTalk Pipeline Component. Start by creating a new project, selecting BizTalk Projects then BizTalk Server Pipeline Component.  You will then be presented with the splash screen. The next step is General Setup, where you will detail the classname, namespace, pipeline and component types, and the implementation language for your Pipeline Component. The options for pipeline type are Receive, Send or Any. Depending on the pipeline type chosen there are different options presented for the component type, matching those available within the BizTalk Pipelines themselves: Receive - Decoder, Disassembling Parser, Validate, Party Resolver, Any. Send -  Encoder, Assembling Serializer, Any. Any - Any. The options for implementation language are C# or VB.Net Next you must set up the UI settings - these are the settings that affect the appearance of the pipeline component within Visual Studio. You must detail the component name, version, description and icon.  Next is the definition of the variables that the pipeline component will use.  The values for these variables will be defined in Visual Studio when creating a pipeline. The options for each variable you require are: Designer Property - The name of the variable. Data Type - String, Boolean, Integer, Long, Short, Schema List, Schema With None Clicking finish now will complete the wizard stage of the creation of your pipeline component. Once the wizard has completed you will be left with a BizTalk Server Pipeline Component project containing a skeleton code file for you to complete.   Within this code file you will mainly be interested in the execute method, which is left mostly empty ready for you to implement your custom pipeline code:          #region IComponent members         /// <summary>         /// Implements IComponent.Execute method.         /// </summary>         /// <param name="pc">Pipeline context</param>         /// <param name="inmsg">Input message</param>         /// <returns>Original input message</returns>         /// <remarks>         /// IComponent.Execute method is used to initiate         /// the processing of the message in this pipeline component.         /// </remarks>         public Microsoft.BizTalk.Message.Interop.IBaseMessage Execute(Microsoft.BizTalk.Component.Interop.IPipelineContext pc, Microsoft.BizTalk.Message.Interop.IBaseMessage inmsg)         {             //             // TODO: implement component logic             //             // this way, it's a passthrough pipeline component             return inmsg;         }         #endregion Once you have implemented your custom code, build and compile your Custom Pipeline Component then add the compiled .dll to C:\Program Files\Microsoft BizTalk Server 2009\Pipeline Components . When creating a new pipeline, in Visual Studio reset the toolbox and the custom pipeline component should appear ready for you to use in your Biztalk Pipeline. Drop the pipeline component into the relevant pipeline stage and configure the component properties (the variables defined in the wizard). You can now deploy and use the pipeline as you would any other custom pipeline.

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• Increasing touch surface (#wp7dev)

  - by Laurent Bugnion

  When you design for Windows Phone 7 (or for any touch device, for that matter, and most especially small screens), you need to be very careful to give enough surface to your users’ fingers. It is easy to miss a touch on such small screens, and that can be horrifyingly frustrating. This is especially true when people are on the move, and trying to hit the control while walking and holding their device in one hand, or when the device is mounted in a car and vibrating with the engine. In my experience, a touch surface should be ideally minimum 60x60 pixels to be easy to activate on the Windows Phone 7 screen (which is, as we know, 800 pixels x 480 pixels). Ideally, I try to make my touch surfaces 80x80 pixels minimum. This causes a few design challenges of course. Using transparent backgrounds However, one thing is helping us tremendously: some surfaces can be made transparent, and yet react to touch. The secret is the following: If you have a panel that has a Null background (i.e. the Background is not set at all), then the empty surface does not react to touch. If however the Background is set to the Transparent color (or any color where the Alpha channel is set to 0), then it will react to touch. Setting a transparent background is easy. For example: <Grid Background="#00000000"> </Grid> or <Grid Background="Transparent"> </Grid> In C#: var grid = new Grid { Background = new SolidColorBrush( Colors.Transparent) }; Using negative margins Having a transparent background reactive to touch is a good start, but in addition, you must make sure that the surface is big enough for my clumsy fingers. One way to achieve that is to increase the transparent, touch-reactive surface, and reposition the element using negative margins. For example, consider the following UI. I changed the transparent background of the HyperlinkButton to Red, in order to visualize the touch surface. In this figure, the Settings HyperlinkButton is 105 pixels x 31 pixels. This is wide enough, but really small in height and easy to miss. To improve this, we can use negative margins, for instance: <HyperlinkButton Content="Settings" HorizontalAlignment="Right" VerticalAlignment="Bottom" Height="60" Margin="0,0,0,-15" /> Notice the usage of negative bottom margin to bring the HyperlinkButton back at the bottom of the main Grid’s first row, where it belongs. And the result is: Notice how the touch surface is much bigger than before. This makes the HyperlinkButton easier to reach, and improves the user experience. With the background set back to normal, the UI looks exactly the same, as it should: In summary: Remember to maximize the touch surface for your controls. Plan your design in consequence by reserving enough room around each control to allow their hit surface to be expanded as shown in this article. Do not cram too many controls in one page. If REALLY needed, use an additional page (or even better: use a Pivot control with multiple pivot items) for the controls that don’t fit on the first one. This should ensure a smoother user experience and improved touch behavior. Happy coding! Laurent   Laurent Bugnion (GalaSoft) Subscribe | Twitter | Facebook | Flickr | LinkedIn

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• How to use a list of values in Excel as filter in a query

  - by Luca Zavarella

  It often happens that a customer provides us with a list of items for which to extract certain information. Imagine, for example, that our clients wish to have the header information of the sales orders only for certain orders. Most likely he will give us a list of items in a column in Excel, or, less probably, a simple text file with the identification code:     As long as the given values ??are at best a dozen, it costs us nothing to copy and paste those values ??in our SSMS and place them in a WHERE clause, using the IN operator, making sure to include the quotes in the case of alphanumeric elements (the database sample is AdventureWorks2008R2): SELECT * FROM Sales.SalesOrderHeader AS SOH WHERE SOH.SalesOrderNumber IN ( 'SO43667' ,'SO43709' ,'SO43726' ,'SO43746' ,'SO43782' ,'SO43796') Clearly, the need to add commas and quotes becomes an hassle when dealing with hundreds of items (which of course has happened to us!). It’d be comfortable to do a simple copy and paste, leaving the items as they are pasted, and make sure the query works fine. We can have this commodity via a User Defined Function, that returns items in a table. Simply we’ll provide the function with an input string parameter containing the pasted items. I give you directly the T-SQL code, where comments are there to clarify what was written: CREATE FUNCTION [dbo].[SplitCRLFList] (@List VARCHAR(MAX)) RETURNS @ParsedList TABLE ( --< Set the item length as your needs Item VARCHAR(255) ) AS BEGIN DECLARE --< Set the item length as your needs @Item VARCHAR(255) ,@Pos BIGINT --< Trim TABs due to indentations SET @List = REPLACE(@List, CHAR(9), '') --< Trim leading and trailing spaces, then add a CR\LF at the end of the list SET @List = LTRIM(RTRIM(@List)) + CHAR(13) + CHAR(10) --< Set the position at the first CR/LF in the list SET @Pos = CHARINDEX(CHAR(13) + CHAR(10), @List, 1) --< If exist other chars other than CR/LFs in the list then... IF REPLACE(@List, CHAR(13) + CHAR(10), '') <> '' BEGIN --< Loop while CR/LFs are over (not found = CHARINDEX returns 0) WHILE @Pos > 0 BEGIN --< Get the heading list chars from the first char to the first CR/LF and trim spaces SET @Item = LTRIM(RTRIM(LEFT(@List, @Pos - 1))) --< If the so calulated item is not empty... IF @Item <> '' BEGIN --< ...insert it in the @ParsedList temporary table INSERT INTO @ParsedList (Item) VALUES (@Item) --(CAST(@Item AS int)) --< Use the appropriate conversion if needed END --< Remove the first item from the list... SET @List = RIGHT(@List, LEN(@List) - @Pos - 1) --< ...and set the position to the next CR/LF SET @Pos = CHARINDEX(CHAR(13) + CHAR(10), @List, 1) --< Repeat this block while the upon loop condition is verified END END RETURN END At this point, having created the UDF, our query is transformed trivially in: SELECT * FROM Sales.SalesOrderHeader AS SOH WHERE SOH.SalesOrderNumber IN ( SELECT Item FROM SplitCRLFList('SO43667 SO43709 SO43726 SO43746 SO43782 SO43796') AS SCL) Convenient, isn’t it? You can find the script DBA_SplitCRLFList.sql here. Bye!!

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• Uninstall, Disable, or Remove Windows 7 Media Center

  - by Mysticgeek

  Although Windows 7 Media Center has improved a lot over previous versions of Windows, but you might want to disable it for different reasons. Here we take a look at a couple of methods to get rid of it. There are a variety of reasons you might want to disable Windows 7 Media Center. Maybe you own a business and don’t want it to run on the machines. Or perhaps you don’t use it at all and just don’t want it around. Turn Off WMC Using Programs and Features Probably the easiest way to get rid of it on all versions of Windows 7 is to open Control Panel and select Programs and Features. This method is similar to disabling Internet Explorer 8 in Windows 7. On the left hand panel click on Turn Windows Features on or off. Scroll down to Media Features and expand the folder. Then Uncheck Windows Media Center… You’ll get a verification message making sure you want to disable it, click Yes. Then the box next to Windows Media Center will be empty…click OK. Wait while WMC is disabled… To complete the process a reboot is required. After getting back from the restart, the WMC icon will be gone and there won’t be any way to launch it. Re-enable WMC If you want to re-enable it, just go back in and recheck it. Again you’ll need to wait while it’s configured, but when it’s done, a restart is not required.   Disable Media Center Using Group Policy Note: This process uses Group Policy Editor which is not available in Home versions of Windows 7. Click on the Start menu and type gpedit.msc into the Search box and hit Enter. Now navigate to User Configuration \ Administrative Templates \ Windows Components \ Windows Media Center. Double-click on Do not allow Windows Media Center to run. Then select the radio button next to Enabled, click OK and close out of Group Policy Editor. Now if a user tries to launch WMC they will get the following message. Conclusion If you’re not a fan of Windows Media Center or want to disable it for whatever reason, the process is simple and there are a couple of ways you can do it. WMC is not included in Starter or Home Basic versions of Windows 7. If you’re new to Windows 7 Media Center, you might want to check out our guide on getting started and setting up live TV. Similar Articles Productive Geek Tips Using Netflix Watchnow in Windows Vista Media Center (Gmedia)Disable Windows Mobility Center in Windows 7 or VistaMake Outlook Faster by Disabling Unnecessary Add-InsSchedule Updates for Windows Media CenterRemove "Map Network Drive" Menu Item from Windows Vista or XP TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Find Downloads and Add-ins for Outlook Recycle ! Find That Elusive Icon with FindIcons Looking for Good Windows Media Player 12 Plug-ins? Find Out the Celebrity You Resemble With FaceDouble Whoa !

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• Subterranean IL: Custom modifiers

  - by Simon Cooper

  In IL, volatile is an instruction prefix used to set a memory barrier at that instruction. However, in C#, volatile is applied to a field to indicate that all accesses on that field should be prefixed with volatile. As I mentioned in my previous post, this means that the field definition needs to store this information somehow, as such a field could be accessed from another assembly. However, IL does not have a concept of a 'volatile field'. How is this information stored? Attributes The standard way of solving this is to apply a VolatileAttribute or similar to the field; this extra metadata notifies the C# compiler that all loads and stores to that field should use the volatile prefix. However, there is a problem with this approach, namely, the .NET C++ compiler. C++ allows methods to be overloaded using properties, like volatile or const, on the parameters; this is perfectly legal C++: public ref class VolatileMethods { void Method(int *i) {} void Method(volatile int *i) {} } If volatile was specified using a custom attribute, then the VolatileMethods class wouldn't be compilable to IL, as there is nothing to differentiate the two methods from each other. This is where custom modifiers come in. Custom modifiers Custom modifiers are similar to custom attributes, but instead of being applied to an IL element separately to its declaration, they are embedded within the field or parameter's type signature itself. The VolatileMethods class would be compiled to the following IL: .class public VolatileMethods { .method public instance void Method(int32* i) {} .method public instance void Method( int32 modreq( [mscorlib]System.Runtime.CompilerServices.IsVolatile)* i) {} } The modreq([mscorlib]System.Runtime.CompilerServices.IsVolatile) is the custom modifier. This adds a TypeDef or TypeRef token to the signature of the field or parameter, and even though they are mostly ignored by the CLR when it's executing the program, this allows methods and fields to be overloaded in ways that wouldn't be allowed using attributes. Because the modifiers are part of the signature, they need to be fully specified when calling such a method in IL: call instance void Method( int32 modreq([mscorlib]System.Runtime.CompilerServices.IsVolatile)*) There are two ways of applying modifiers; modreq specifies required modifiers (like IsVolatile), and modopt specifies optional modifiers that can be ignored by compilers (like IsLong or IsConst). The type specified as the modifier argument are simple placeholders; if you have a look at the definitions of IsVolatile and IsLong they are completely empty. They exist solely to be referenced by a modifier. Custom modifiers are used extensively by the C++ compiler to specify concepts that aren't expressible in IL, but still need to be taken into account when calling method overloads. C++ and C# That's all very well and good, but how does this affect C#? Well, the C++ compiler uses modreq(IsVolatile) to specify volatility on both method parameters and fields, as it would be slightly odd to have the same concept represented using a modifier or attribute depending on what it was applied to. Once you've compiled your C++ project, it can then be referenced and used from C#, so the C# compiler has to recognise the modreq(IsVolatile) custom modifier applied to fields, and vice versa. So, even though you can't overload fields or parameters with volatile using C#, volatile needs to be expressed using a custom modifier rather than an attribute to guarentee correct interoperability and behaviour with any C++ dlls that happen to come along. Next up: a closer look at attributes, and how certain attributes compile in unexpected ways.

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• Love and Hate Outlook autocomplete, Outlook 2010/Exchange 2010

  - by Kay Sellenrode

  I think that almost every Exchange admin can concur with me that the Outlook autocomplete cache is one of those things you love but at the same time also hate. Users mostly love this function, except when it fails.Luckily since Outlook 2010 things got a little better and we got rid of the dreaded nk2 files.Outlook 2010 now includes a folder named "Suggested Contacts", all users you send an email to and that don't already have an contact object are saved in this suggested contacts folder.A lot of people thought this folder is also the source for the autocomplete cache, which would make it somewhat easy to manage, I wish the solution was that easy.Badly enough separate from the suggested contacts, outlook still maintains a cache for the autocomplete function. Let us say you run in to the following situation: John works for company A and is a popular contact for almost everyone in your organization.Now John quit his job at Company A and moved to Company B.Luckily John maintains your company as customer, but his email address is now changed from companyA.com to companyB.comSince you don't want to do any business with Company A anymore, you want to make sure none of your users accidentally mail to his old address.Now this is where the real fun starts, cause almost all of your 1000 users have mailed at least once with John.Resulting in the fact that every user has John most probably listed in their autocomplete cache.  I have run into sort like situations multiple times with several customers, which is always a pain.And of course this blog post is the result of one of those issues once again.I knew that with the Suggested contacts we could do more than previously, but still never spent time on it before.But today I thought lets nail this now and forever!!  Ok let's start of that things are different for every combination of outlook and exchange.I explain the procedure for Exchange 2010 SP1+ in combination with Outlook 2010.At first we want to get rid of all contact objects that contain [email protected] do this we need to be assigned to the RBAC role "Mailbox Import Export", which can be done through the Exchange Control panel.In my test environment I assigned this role to the Organization admins, but in real life you might want to add it to a custom role. Open the Exchange control panel by logging in to the ecp url, in my case https://ITFEX.itf.local/ECP, and make sure you selected your organization as management scope.Browse to Roles & Auditing, and open the properties for the organization management role group.click on the Add button to add a new role to the Organization Management role group, select the Mailbox Import Export role and click on add and OK to add it to the role.  Once you have assigned that role to your account you can open the Exchange Management Shell and execute the following command: Get-mailbox –resultsize unlimited | search-mailbox –targetmailbox "your.account" –targetfolder searchanddelete –loglevel full –logonly –searchquery "kind:contact AND [email protected]" This command will create a list with all mailboxes and any contacts that were found with an email address that contains [email protected], this list is then posted in the mailbox you specified at your.account in the folder searchanddelete.Now examine the report that was created and posted in the mailbox to see if it matches what you think it should match.My results looked like this:  When you're confident that the search includes all references and no false positives you can execute almost the same command, but this time with an delete action instead of the logonly. Get-mailbox –resultsize unlimited | search-mailbox –targetmailbox "your.account" –targetfolder searchanddelete –loglevel full –DeleteContent –searchquery "kind:contact AND [email protected]" Now most people would think this would remove the contact object from the suggested contacts, resulting in a removal from the autocomplete list.Sad but not true, to clean up the autocomplete list start Outlook with the command: "outlook /cleanautocompletecache" This will result in an empty cache, but luckily this is rebuild based on the suggested contacts, which now doesn't include the [email protected] contact anymore.

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• App Stores&ndash;In All Things, Its Quality Over Quantity

  - by D'Arcy Lussier

  Everybody has an opinion about Windows 8. People love it, people hate it, people are meh about it, people are apparently buying it from Microsoft stores in NYC as if it was water before a natural disaster…if there’s one thing that Microsoft product launches do well, its the ability to bring out strong emotional responses. Over at eweek.com, Don Reisinger wrote about 5 good and bad things about Windows 8. Yes, another opinion piece on WIndows 8. I figured since this one had good and bad it might be worthwhile to read. I then came across #10 on his list, and figured “What the hell…might as well post a bit of a rant on Windows 8 myself!” Here’s #10: 10. Bad: Too few apps Unfortunately, Microsoft wasn’t able to get too many developers to start producing applications for its Windows 8 Store. Microsoft hasn’t yet released official numbers, but some have said that the marketplace has less than 8,000 programs. Considering Apple’s App Store has 100 times that, it’s about time Microsoft starts leaning on developers to get more programs into its store. Believe me, Microsoft *has* been leaning on developers to get apps into the store. I’ve been asked at least 5 or 6 times from 5 or 6 different friends at Microsoft about whether I was going to write a Windows 8 app. I think Microsoft felt they had to try and address the number of apps available in their marketplace, since some people (like Don) would draw comparisons to the number of apps in the Apple marketplace. I feel for Microsoft in this, since the number of apps in a marketplace are an empty stat. Quality of Quantity I have an iPad that my family (wife, 10yo daughter, 3yo daughter) use. We all have our own apps installed on it. In addition, my wife has an iPhone 4S that she also installs apps on. As someone who gets asked by his kids often whether they can buy/download an app, the vast majority of the vast catalogue of iOS marketplace apps are crap! Do you realize how many “free” games are out there, only to really be not-free because you have to purchase in-game content to make the game actually playable? And how about searching – with such a vast array of apps and such high numbers of craptastic ones, trying to find something is incredibly difficult and can be frustrating. I would rather see that Microsoft has 8000 high quality apps in their store at launch, instead of 800000 that were mostly junk. Too Few Apps?! And seriously, 8000 is not a small number. How many iOS apps have I actually bought between the iPad and iPhone? I’ll be generous and say 30…heck, let’s round it up to 40. It’s not like I have 10,000 apps installed on my iPad, nor will that ever happen! So if people have, at the *launch* of a new platform ecosystem, EIGHT THOUSAND apps to choose from, I don’t see that as a fail at all! It should be noted that most of the most common apps (Netflix, Skype, etc.) are available for Windows 8 at launch – I guess I’ll have to wait a few weeks for My Pony Ranch and all its clones to start showing up; pity. Let’s Check Back in a Year So look, let’s check back in a year’s time and see what the app store looks like. My hope is that Microsoft doesn’t continue to push quantity over quality. Even knowing the optics that # of apps in the store carries and the pressure to catch Apple and Android marketplaces, I hope Microsoft avoids the scenario where there’s a good percentage of apps in the Windows Store that are utter rubbish and finding the gems will be cumbersome. But if that happens, we can thank guys like Dan who raised the false issue of app count at the launch for it.

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• wifi hardware switch doesn't work on a Dell 1018

  - by user42566

  I have a problem with my Dell 1018 Inspiron. I can't switch the wifi on, through the key on the keyboard. I think it's a driver problem since Ubuntu 11.10. This are the versions i tried: Ubuntu 10.04 / 10.10 It's possible to install the driver by hand: sudo add-apt-repository ppa:lexical/hwe-wireless sudo apt-get update sudo apt-get install rtl8192ce-dkms Ubuntu 11.04 It works out of the box Ubuntu 11.10 / 12.04 I haven’t found any solution for these versions. The "ppa:lexical/hwe-wireless" doesn't work for these versions. It says Can not find package rtl8192ce-dkms. The window of additional drivers is empty. So I can't install the driver. The wired network works good. Here is some information: 0: dell-wifi: Wireless LAN Soft blocked: no Hard blocked: no 1: phy0: Wireless LAN Soft blocked: no Hard blocked: yes sudo lshw -class network *-network description: Ethernet interface product: RTL8101E/RTL8102E PCI Express Fast Ethernet controller vendor: Realtek Semiconductor Co., Ltd. physical id: 0 bus info: pci@0000:05:00.0 logical name: eth0 version: 05 serial: 5c:26:0a:0d:20:10 size: 10Mbit/s capacity: 100Mbit/s width: 64 bits clock: 33MHz capabilities: pm msi pciexpress msix vpd bus_master cap_list ethernet physical tp mii 10bt 10bt-fd 100bt 100bt-fd autonegotiation configuration: autonegotiation=on broadcast=yes driver=r8169 driverversion=2.3LK-NAPI duplex=half firmware=rtl_nic/rtl8105e-1.fw latency=0 link=no multicast=yes port=MII speed=10Mbit/s resources: irq:43 ioport:2000(size=256) memory:f0f2c000-f0f2cfff memory:f0f18000-f0f1bfff *-network description: Wireless interface product: RTL8188CE 802.11b/g/n WiFi Adapter vendor: Realtek Semiconductor Co., Ltd. physical id: 0 bus info: pci@0000:07:00.0 logical name: wlan0 version: 01 serial: 70:f1:a1:fe:15:bd width: 64 bits clock: 33MHz capabilities: pm msi pciexpress bus_master cap_list ethernet physical wireless configuration: broadcast=yes driver=rtl8192ce driverversion=3.2.0-22-generic-pae firmware=N/A ip=192.168.1.76 latency=0 link=yes multicast=yes wireless=IEEE 802.11bgn resources: irq:17 ioport:3000(size=256) memory:f0100000-f0103fff mark@mark-Inspiron-1018:~$ mark@mark-Inspiron-1018:~$ sudo lspci -nn 00:00.0 Host bridge [0600]: Intel Corporation N10 Family DMI Bridge [8086:a010] 00:02.0 VGA compatible controller [0300]: Intel Corporation N10 Family Integrated Graphics Controller [8086:a011] 00:02.1 Display controller [0380]: Intel Corporation N10 Family Integrated Graphics Controller [8086:a012] 00:1b.0 Audio device [0403]: Intel Corporation N10/ICH 7 Family High Definition Audio Controller [8086:27d8] (rev 02) 00:1c.0 PCI bridge [0604]: Intel Corporation N10/ICH 7 Family PCI Express Port 1 [8086:27d0] (rev 02) 00:1c.1 PCI bridge [0604]: Intel Corporation N10/ICH 7 Family PCI Express Port 2 [8086:27d2] (rev 02) 00:1d.0 USB controller [0c03]: Intel Corporation N10/ICH 7 Family USB UHCI Controller #1 [8086:27c8] (rev 02) 00:1d.1 USB controller [0c03]: Intel Corporation N10/ICH 7 Family USB UHCI Controller #2 [8086:27c9] (rev 02) 00:1d.2 USB controller [0c03]: Intel Corporation N10/ICH 7 Family USB UHCI Controller #3 [8086:27ca] (rev 02) 00:1d.3 USB controller [0c03]: Intel Corporation N10/ICH 7 Family USB UHCI Controller #4 [8086:27cb] (rev 02) 00:1d.7 USB controller [0c03]: Intel Corporation N10/ICH 7 Family USB2 EHCI Controller [8086:27cc] (rev 02) 00:1e.0 PCI bridge [0604]: Intel Corporation 82801 Mobile PCI Bridge [8086:2448] (rev e2) 00:1f.0 ISA bridge [0601]: Intel Corporation NM10 Family LPC Controller [8086:27bc] (rev 02) 00:1f.2 SATA controller [0106]: Intel Corporation N10/ICH7 Family SATA Controller [AHCI mode] [8086:27c1] (rev 02) 00:1f.3 SMBus [0c05]: Intel Corporation N10/ICH 7 Family SMBus Controller [8086:27da] (rev 02) 05:00.0 Ethernet controller [0200]: Realtek Semiconductor Co., Ltd. RTL8101E/RTL8102E PCI Express Fast Ethernet controller [10ec:8136] (rev 05) 07:00.0 Network controller [0280]: Realtek Semiconductor Co., Ltd. RTL8188CE 802.11b/g/n WiFi Adapter [10ec:8176] (rev 01) mark@mark-Inspiron-1018:~$ mark@mark-Inspiron-1018:~$ lsusb Bus 001 Device 001: ID 1d6b:0002 Linux Foundation 2.0 root hub Bus 002 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 003 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 004 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 005 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 001 Device 002: ID 174f:1127 Syntek mark@mark-Inspiron-1018:~$

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• Building ATLAS (and later Octave w/ ATLAS)

  - by David Parks

  I'm trying to set up ATLAS (so I can later compile octave with ATLAS support). If I'm correct, I still need to build this manually due to the environment specific optimizations. I do see a package for ATLAS, but it looks like it's using the cross platform generic build options (e.g. "it'll be slow"). So, running the configure script as described in the docs seems to go poorly. As a java developer I never do well at making heads or tails of errors in these build processes. Am I missing dependencies (if so is there any documentation on what I need)? allusers@vbubuntu:~/Downloads/atlas3.10.1/build_vbubuntu$ ../configure -b 64 -D c -DPentiumCPS=3000 --with-netlib-lapack-tarfile=/home/allusers/Downloads/lapack-3.5.0.tgz make: `xconfig' is up to date. ./xconfig -d s /home/allusers/Downloads/atlas3.10.1/build_vbubuntu/../ -d b /home/allusers/Downloads/atlas3.10.1/build_vbubuntu -b 64 -D c -DPentiumCPS=3000 -Si lapackref 1 OS configured as Linux (1) Assembly configured as GAS_x8664 (2) Vector ISA Extension configured as SSE3 (6,448) ERROR: enum fam=3, chip=2, mach=0 make[3]: *** [atlas_run] Error 44 make[2]: *** [IRunArchInfo_x86] Error 2 Architecture configured as Corei1 (25) ERROR: enum fam=3, chip=2, mach=0 make[3]: *** [atlas_run] Error 44 make[2]: *** [IRunArchInfo_x86] Error 2 Clock rate configured as 2350Mhz ERROR: enum fam=3, chip=2, mach=0 make[3]: *** [atlas_run] Error 44 make[2]: *** [IRunArchInfo_x86] Error 2 Maximum number of threads configured as 4 Parallel make command configured as '$(MAKE) -j 4' ERROR: enum fam=3, chip=2, mach=0 make[3]: *** [atlas_run] Error 44 make[2]: *** [IRunArchInfo_x86] Error 2 Cannot detect CPU throttling. rm -f config1.out make atlas_run atldir=/home/allusers/Downloads/atlas3.10.1/build_vbubuntu exe=xprobe_comp redir=config1.out \ args="-v 0 -o atlconf.txt -O 1 -A 25 -Si nof77 0 -V 448 -b 64 -d b /home/allusers/Downloads/atlas3.10.1/build_vbubuntu" make[1]: Entering directory `/home/allusers/Downloads/atlas3.10.1/build_vbubuntu' cd /home/allusers/Downloads/atlas3.10.1/build_vbubuntu ; ./xprobe_comp -v 0 -o atlconf.txt -O 1 -A 25 -Si nof77 0 -V 448 -b 64 -d b /home/allusers/Downloads/atlas3.10.1/build_vbubuntu > config1.out make[2]: gfortran: Command not found make[2]: *** [IRunF77Comp] Error 127 make[2]: g77: Command not found make[2]: *** [IRunF77Comp] Error 127 make[2]: f77: Command not found make[2]: *** [IRunF77Comp] Error 127 Unable to find usable compiler for F77; abortingMake sure compilers are in your path, and specify good compilers to configure (see INSTALL.txt or 'configure --help' for details)make[1]: *** [atlas_run] Error 8 make[1]: Leaving directory `/home/allusers/Downloads/atlas3.10.1/build_vbubuntu' make: *** [IRun_comp] Error 2 ERROR 512 IN SYSCMND: 'make IRun_comp args="-v 0 -o atlconf.txt -O 1 -A 25 -Si nof77 0 -V 448 -b 64"' mkdir src bin tune interfaces mkdir: cannot create directory ‘src’: File exists mkdir: cannot create directory ‘bin’: File exists mkdir: cannot create directory ‘tune’: File exists mkdir: cannot create directory ‘interfaces’: File exists make: *** [make_subdirs] Error 1 make -f Make.top startup make[1]: Entering directory `/home/allusers/Downloads/atlas3.10.1/build_vbubuntu' Make.top:1: Make.inc: No such file or directory Make.top:325: warning: overriding commands for target `/AtlasTest' Make.top:76: warning: ignoring old commands for target `/AtlasTest' make[1]: *** No rule to make target `Make.inc'. Stop. make[1]: Leaving directory `/home/allusers/Downloads/atlas3.10.1/build_vbubuntu' make: *** [startup] Error 2 mv: cannot move ‘lapack-3.5.0’ to ‘../reference/lapack-3.5.0’: Directory not empty mv: cannot stat ‘lib/Makefile’: No such file or directory ../configure: 450: ../configure: cannot create lib/Makefile: Directory nonexistent ../configure: 451: ../configure: cannot create lib/Makefile: Directory nonexistent ../configure: 452: ../configure: cannot create lib/Makefile: Directory nonexistent ../configure: 453: ../configure: cannot create lib/Makefile: Directory nonexistent ../configure: 509: ../configure: cannot create lib/Makefile: Directory nonexistent DONE configure

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• Azure Mobile Services: lessons learned

  - by svdoever

  When I first started using Azure Mobile Services I thought of it as a nice way to: authenticate my users - login using Twitter, Google, Facebook, Windows Live create tables, and use the client code to create the columns in the table because that is not possible in the Azure Mobile Services UI run some Javascript code on the table crud actions (Insert, Update, Delete, Read) schedule a Javascript to run any 15 or more minutes I had no idea of the magic that was happening inside… where is the data stored? Is it a kind of big table, are relationships between tables possible? those Javascripts on the table crud actions, is that interpreted, what is that exactly? After working for some time with Azure Mobile Services I became a lot wiser: Those tables are just normal tables in an Azure SQL Server 2012 Creating the table columns through client code sucks, at least from my Javascript code, because the columns are deducted from the sent JSON data, and a datetime field is sent as string in JSON, so a string type column is created instead of a datetime column You can connect with SQL Management Studio to the Azure SQL Server, and although you can’t manage your columns through the SQL Management Studio UI, it is possible to just run SQL scripts to drop and create tables and indices When you create a table through SQL script, add the table with the same name in the Azure Mobile Services UI to hook it up and be able to access the table through the provided abstraction layer You can also go to the SQL Database through the Azure Mobile Services UI, and from there get in a web based SQL management studio where you can create columns and manage your data The table crud scripts and the scheduler scripts are full blown node.js scripts, introducing a lot of power with great performance The web based script editor is really powerful, I do most of my editing currently in the editor which has syntax highlighting and code completing. While editing the code JsHint is used for script validation. The documentation on Azure Mobile Services is… suboptimal. It is such a pity that there is no way to comment on it so the community could fill in the missing holes, like which node modules are already loaded, and which modules are available on Azure Mobile Services. Soon I was hacking away on Azure Mobile Services, creating my own database tables through script, and abusing the read script of an empty table named query to implement my own set of “services”. The latest updates to Azure Mobile Services described in the following posts added some great new features like creating web API’s, use shared code from your scripts, command line tools for managing Azure Mobile Services (upload and download scripts for example), support for node modules and git support: http://weblogs.asp.net/scottgu/archive/2013/06/14/windows-azure-major-updates-for-mobile-backend-development.aspx http://blogs.msdn.com/b/carlosfigueira/archive/2013/06/14/custom-apis-in-azure-mobile-services.aspx http://blogs.msdn.com/b/carlosfigueira/archive/2013/06/19/custom-api-in-azure-mobile-services-client-sdks.aspx In the mean time I rewrote all my “service-like” table scripts to API scripts, which works like a breeze. Bad thing with the current state of Azure Mobile Services is that the git support is not working if you are a co-administrator of your Azure subscription, and not and administrator (as in my case). Another bad thing is that Cross Origin Request Sharing (CORS) is not supported for the API yet, so no go yet from the browser client for API’s, which is my case. See http://social.msdn.microsoft.com/Forums/windowsazure/en-US/2b79c5ea-d187-4c2b-823a-3f3e0559829d/known-limitations-for-source-control-and-custom-api-features for more on these and other limitations. In his talk at Build 2013 Josh Twist showed that there is a work-around for accessing shared script code from the table scripts as well (another limitation mentioned in the post above). I could not find that code in the Votabl2 code example from the presentation at https://github.com/joshtwist/votabl2, but we can grab it from the presentation when it comes online on Channel9. By the way: you can always express your needs and ideas at http://mobileservices.uservoice.com, that’s the place they are listening to (I hope!).

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• Sharing configuration settings between Windows Azure roles

  - by theo.spears

  If you are working on a medium-large Windows Azure project it's likely it will involve more than one role, for example separate web and worker roles. Unfortunately although all the windows azure configuration settings are stored in a single cscfg file, there is no way to share configuration settings between multiple roles. This means you have to duplicate common settings like connection strings across all your roles. There is an open Connect issue about this topic, but Microsoft have not said when they will fix it. In the mean time I've put together a dirty dirty hack cunning workaround that creates a fake role containing your shared configuration settings, and copies it to all roles as part of the build process. Here's how you set it up: 1. Download the zip file attached to this post, and unzip it into the folder containing your Azure project (not your solution folder). 2. Edit your csdef and cscfg files to include the placeholder project ServiceDefinition.csdef<?xml version="1.0" encoding="utf-8"?> <ServiceDefinition name="AzureSpendNotifier" http://schemas.microsoft.com/ServiceHosting/2008/10/ServiceDefinition%22"http://schemas.microsoft.com/ServiceHosting/2008/10/ServiceDefinition"> <WorkerRole name="GLOBAL"> <ConfigurationSettings> <Setting name="ExampleSetting" /> </ConfigurationSettings> </WorkerRole> <WorkerRole name="MyWorker"> <ConfigurationSettings> </ConfigurationSettings> </WorkerRole> <WebRole name="MyWeb"> <Sites> <Site name="Web"> <Bindings> <Binding name="WebEndpoint" endpointName="WebEndpoint" /> </Bindings> </Site> </Sites> <ConfigurationSettings> </ConfigurationSettings> </WebRole> </ServiceDefinition> ServiceConfiguration.cscfg<?xml version="1.0" encoding="utf-8"?> <ServiceConfiguration serviceName="AzureSpendNotifier" xmlns=http://schemas.microsoft.com/ServiceHosting/2008/10/ServiceConfiguration osFamily="1" osVersion="*"> <Role name="GLOBAL"> <ConfigurationSettings> <Setting name="ExampleSetting" value="Hello World" /> </ConfigurationSettings> <Instances count="1" /> </Role> <Role name="MyWorker"> <Instances count="1" /> <ConfigurationSettings> </ConfigurationSettings> </Role> <Role name="MyWeb"> <Instances count="1" /> <ConfigurationSettings> </ConfigurationSettings> </Role> </ServiceConfiguration> It is important that all your roles contain a ConfigurationSettings entry in both cscfg and csdef files, even if it's empty- otherwise the shared configuration settings will not be inserted. 3. Open your azure deployment (.ccproj) project in notepad, and add the highlighted line below: ... <Import Project="$(CloudExtensionsDir)Microsoft.CloudService.targets" /> <Import Project="globalsettings/globalsettings.targets" /> </Project> It is important you add this below the Microsoft.CloudService.targets import line, as it replaces some of the rules defined in that file. Visual studio will prompt you to reload the project, say yes. At this point you will have a new Azure role called 'GLOBAL' with settings you can edit through the visual studio properties panel as normal. This role will never be deployed, but any settings you add to it will be copied to all your other roles when deployed or tested locally within visual studio.

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• Project Euler 17: (Iron)Python

  - by Ben Griswold

  In my attempt to learn (Iron)Python out in the open, here’s my solution for Project Euler Problem 17.  As always, any feedback is welcome. # Euler 17 # http://projecteuler.net/index.php?section=problems&id=17 # If the numbers 1 to 5 are written out in words: # one, two, three, four, five, then there are # 3 + 3 + 5 + 4 + 4 = 19 letters used in total. # If all the numbers from 1 to 1000 (one thousand) # inclusive were written out in words, how many letters # would be used? # # NOTE: Do not count spaces or hyphens. For example, 342 # (three hundred and forty-two) contains 23 letters and # 115 (one hundred and fifteen) contains 20 letters. The # use of "and" when writing out numbers is in compliance # with British usage. import time start = time.time() def to_word(n): h = { 1 : "one", 2 : "two", 3 : "three", 4 : "four", 5 : "five", 6 : "six", 7 : "seven", 8 : "eight", 9 : "nine", 10 : "ten", 11 : "eleven", 12 : "twelve", 13 : "thirteen", 14 : "fourteen", 15 : "fifteen", 16 : "sixteen", 17 : "seventeen", 18 : "eighteen", 19 : "nineteen", 20 : "twenty", 30 : "thirty", 40 : "forty", 50 : "fifty", 60 : "sixty", 70 : "seventy", 80 : "eighty", 90 : "ninety", 100 : "hundred", 1000 : "thousand" } word = "" # Reverse the numbers so position (ones, tens, # hundreds,...) can be easily determined a = [int(x) for x in str(n)[::-1]] # Thousands position if (len(a) == 4 and a[3] != 0): # This can only be one thousand based # on the problem/method constraints word = h[a[3]] + " thousand " # Hundreds position if (len(a) >= 3 and a[2] != 0): word += h[a[2]] + " hundred" # Add "and" string if the tens or ones # position is occupied with a non-zero value. # Note: routine is broken up this way for [my] clarity. if (len(a) >= 2 and a[1] != 0): # catch 10 - 99 word += " and" elif len(a) >= 1 and a[0] != 0: # catch 1 - 9 word += " and" # Tens and ones position tens_position_value = 99 if (len(a) >= 2 and a[1] != 0): # Calculate the tens position value per the # first and second element in array # e.g. (8 * 10) + 1 = 81 tens_position_value = int(a[1]) * 10 + a[0] if tens_position_value <= 20: # If the tens position value is 20 or less # there's an entry in the hash. Use it and there's # no need to consider the ones position word += " " + h[tens_position_value] else: # Determine the tens position word by # dividing by 10 first. E.g. 8 * 10 = h[80] # We will pick up the ones position word later in # the next part of the routine word += " " + h[(a[1] * 10)] if (len(a) >= 1 and a[0] != 0 and tens_position_value > 20): # Deal with ones position where tens position is # greater than 20 or we have a single digit number word += " " + h[a[0]] # Trim the empty spaces off both ends of the string return word.replace(" ","") def to_word_length(n): return len(to_word(n)) print sum([to_word_length(i) for i in xrange(1,1001)]) print "Elapsed Time:", (time.time() - start) * 1000, "millisecs" a=raw_input('Press return to continue')

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• Improving the running time of Breadth First Search and Adjacency List creation

  - by user45957

  We are given an array of integers where all elements are between 0-9. have to start from the 1st position and reach end in minimum no of moves such that we can from an index i move 1 position back and forward i.e i-1 and i+1 and jump to any index having the same value as index i. Time Limit : 1 second Max input size : 100000 I have tried to solve this problem use a single source shortest path approach using Breadth First Search and though BFS itself is O(V+E) and runs in time the adjacency list creation takes O(n2) time and therefore overall complexity becomes O(n2). is there any way i can decrease the time complexity of adjacency list creation? or is there a better and more efficient way of solving the problem? int main(){ vector<int> v; string str; vector<int> sets[10]; cin>>str; int in; for(int i=0;i<str.length();i++){ in=str[i]-'0'; v.push_back(in); sets[in].push_back(i); } int n=v.size(); if(n==1){ cout<<"0\n"; return 0; } if(v[0]==v[n-1]){ cout<<"1\n"; return 0; } vector<int> adj[100001]; for(int i=0;i<10;i++){ for(int j=0;j<sets[i].size();j++){ if(sets[i][j]>0) adj[sets[i][j]].push_back(sets[i][j]-1); if(sets[i][j]<n-1) adj[sets[i][j]].push_back(sets[i][j]+1); for(int k=j+1;k<sets[i].size();k++){ if(abs(sets[i][j]-sets[i][k])!=1){ adj[sets[i][j]].push_back(sets[i][k]); adj[sets[i][k]].push_back(sets[i][j]); } } } } queue<int> q; q.push(0); int dist[100001]; bool visited[100001]={false}; dist[0]=0; visited[0]=true; int c=0; while(!q.empty()){ int dq=q.front(); q.pop(); c++; for(int i=0;i<adj[dq].size();i++){ if(visited[adj[dq][i]]==false){ dist[adj[dq][i]]=dist[dq]+1; visited[adj[dq][i]]=true; q.push(adj[dq][i]); } } } cout<<dist[n-1]<<"\n"; return 0; }

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• Best Practices - Core allocation

  - by jsavit

  This post is one of a series of "best practices" notes for Oracle VM Server for SPARC (also called Logical Domains) Introduction SPARC T-series servers currently have up to 4 CPU sockets, each of which has up to 8 or (on SPARC T3) 16 CPU cores, while each CPU core has 8 threads, for a maximum of 512 dispatchable CPUs. The defining feature of Oracle VM Server for SPARC is that each domain is assigned CPU threads or cores for its exclusive use. This avoids the overhead of software-based time-slicing and emulation (or binary rewriting) of system state-changing privileged instructions used in traditional hypervisors. To create a domain, administrators specify either the number of CPU threads or cores that the domain will own, as well as its memory and I/O resources. When CPU resources are assigned at the individual thread level, the logical domains constraint manager attempts to assign threads from the same cores to a domain, and avoid "split core" situations where the same CPU core is used by multiple domains. Sometimes this is unavoidable, especially when domains are allocated and deallocated CPUs in small increments. Why split cores can matter Split core allocations can silenty reduce performance because multiple domains with different address spaces and memory contents are sharing the core's Level 1 cache (L1$). This is called false cache sharing since even identical memory addresses from different domains must point to different locations in RAM. The effect of this is increased contention for the cache, and higher memory latency for each domain using that core. The degree of performance impact can be widely variable. For applications with very small memory working sets, and with I/O bound or low-CPU utilization workloads, it may not matter at all: all machines wait for work at the same speed. If the domains have substantial workloads, or are critical to performance then this can have an important impact: This blog entry was inspired by a customer issue in which one CPU core was split among 3 domains, one of which was the control and service domain. The reported problem was increased I/O latency in guest domains, but the root cause might be higher latency servicing the I/O requests due to the control domain being slowed down. What to do about it Split core situations are easily avoided. In most cases the logical domain constraint manager will avoid it without any administrative action, but it can be entirely prevented by doing one of the several actions: Assign virtual CPUs in multiples of 8 - the number of threads per core. For example: ldm set-vcpu 8 mydomain or ldm add-vcpu 24 mydomain. Each domain will then be allocated on a core boundary. Use the whole core constraint when assigning CPU resources. This allocates CPUs in increments of entire cores instead of virtual CPU threads. The equivalent of the above commands would be ldm set-core 1 mydomain or ldm add-core 3 mydomain. Older syntax does the same thing by adding the -c flag to the add-vcpu, rm-vcpu and set-vcpu commands, but the new syntax is recommended. When whole core allocation is used an attempt to add cores to a domain fails if there aren't enough completely empty cores to satisfy the request. See https://blogs.oracle.com/sharakan/entry/oracle_vm_server_for_sparc4 for an excellent article on this topic by Eric Sharakan. Don't obsess: - if the workloads have minimal CPU requirements and don't need anywhere near a full CPU core, then don't worry about it. If you have low utilization workloads being consolidated from older machines onto a current T-series, then there's no need to worry about this or to assign an entire core to domains that will never use that much capacity. In any case, make sure the most important domains have their own CPU cores, in particular the control domain and any I/O or service domain, and of course any important guests. Summary Split core CPU allocation to domains can potentially have an impact on performance, but the logical domains manager tends to prevent this situation, and it can be completely and simply avoided by allocating virtual CPUs on core boundaries.

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• Node.js Lockstep Multiplayer Architecture

  - by Wakaka

  Background I'm using the lockstep model for a multiplayer Node.js/Socket.IO game in a client-server architecture. User input (mouse or keypress) is parsed into commands like 'attack' and 'move' on the client, which are sent to the server and scheduled to be executed on a certain tick. This is in contrast to sending state data to clients, which I don't wish to use due to bandwidth issues. Each tick, the server will send the list of commands on that tick (possibly empty) to each client. The server and all clients will then process the commands and simulate that tick in exactly the same way. With Node.js this is actually quite simple due to possibility of code sharing between server and client. I'll just put the deterministic simulator in the /shared folder which can be run by both server and client. The server simulation is required so that there is an authoritative version of the simulation which clients cannot alter. Problem Now, the game has many entity classes, like Unit, Item, Tree etc. Entities are created in the simulator. However, for each class, it has some methods that are shared and some that are client-specific. For instance, the Unit class has addHp method which is shared. It also has methods like getSprite (gets the image of the entity), isVisible (checks if unit can be seen by the client), onDeathInClient (does a bunch of stuff when it dies only on the client like adding announcements) and isMyUnit (quick function to check if the client owns the unit). Up till now, I have been piling all the client functions into the shared Unit class, and adding a this.game.isServer() check when necessary. For instance, when the unit dies, it will call if (!this.game.isServer()) { this.onDeathInClient(); }. This approach has worked pretty fine so far, in terms of functionality. But as the codebase grew bigger, this style of coding seems a little strange. Firstly, the client code is clearly not shared, and yet is placed under the /shared folder. Secondly, client-specific variables for each entity are also instantiated on the server entity (like unit.sprite) and can run into problems when the server cannot instantiate the variable (it doesn't have Image class like on browsers). So my question is, is there a better way to organize the client code, or is this a common way of doing things for lockstep multiplayer games? I can think of a possible workaround, but it does have its own problems. Possible workaround (with problems) I could use Javascript mixins that are only added when in a browser. Thus, in the /shared/unit.js file in the /shared folder, I would have this code at the end: if (typeof exports !== 'undefined') module.exports = Unit; else mixin(Unit, LocalUnit); Then I would have /client/localunit.js store an object LocalUnit of client-side methods for Unit. Now, I already have a publish-subscribe system in place for events in the simulator. To remove the this.game.isServer() checks, I could publish entity-specific events whenever I want the client to do something. For instance, I would do this.publish('Death') in /shared/unit.js and do this.subscribe('Death', this.onDeathInClient) in /client/localunit.js. But this would make the simulator's event listeners list on the server and the client different. Now if I want to clear all subscribed events only from the shared simulator, I can't. Of course, it is possible to create two event subscription systems - one client-specific and one shared - but now the publish() method would have to do if (!this.game.isServer()) { this.publishOnClient(event); }. All in all, the workaround off the top of my head seems pretty complicated for something as simple as separating the client and shared code. Thus, I wonder if there is an established and simpler method for better code organization, hopefully specific to Node.js games.

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• Question on the implementation of my Entity System

  - by miguel.martin

  I am currently creating an Entity System, in C++, it is almost completed (I have all the code there, I just have to add a few things and test it). The only thing is, I can't figure out how to implement some features. This Entity System is based off a bit from the Artemis framework, however it is different. I'm not sure if I'll be able to type this out the way my head processing it. I'm going to basically ask whether I should do something over something else. Okay, now I'll give a little detail on my Entity System itself. Here are the basic classes that my Entity System uses to actually work: Entity - An Id (and some methods to add/remove/get/etc Components) Component - An empty abstract class ComponentManager - Manages ALL components for ALL entities within a Scene EntitySystem - Processes entities with specific components Aspect - The class that is used to help determine what Components an Entity must contain so a specific EntitySystem can process it EntitySystemManager - Manages all EntitySystems within a Scene EntityManager - Manages entities (i.e. holds all Entities, used to determine whether an Entity has been changed, enables/disables them, etc.) EntityFactory - Creates (and destroys) entities and assigns an ID to them Scene - Contains an EntityManager, EntityFactory, EntitySystemManager and ComponentManager. Has functions to update and initialise the scene. Now in order for an EntitySystem to efficiently know when to check if an Entity is valid for processing (so I can add it to a specific EntitySystem), it must recieve a message from the EntityManager (after a call of activate(Entity& e)). Similarly the EntityManager must know when an Entity is destroyed from the EntityFactory in the Scene, and also the ComponentManager must know when an Entity is created AND destroyed. I do have a Listener/Observer pattern implemented at the moment, but with this pattern I may remove a Listener (which is this case is dependent on the method being called). I mainly have this implemented for specific things related to a game, i.e. Teams, Tagging of entities, etc. So... I was thinking maybe I should call a private method (using friend classes) to send out when an Entity has been activated, deleted, etc. i.e. taken from my EntityFactory void EntityFactory::killEntity(Entity& e) { // if the entity doesn't exsist in the entity manager within the scene if(!getScene()->getEntityManager().doesExsist(e)) { return; // go back to the caller! (should throw an exception or something..) } // tell the ComponentManager and the EntityManager that we killed an Entity getScene()->getComponentManager().doOnEntityWillDie(e); getScene()->getEntityManager().doOnEntityWillDie(e); // notify the listners for(Mouth::iterator i = getMouth().begin(); i != getMouth().end(); ++i) { (*i)->onEntityWillDie(*this, e); } _idPool.addId(e.getId()); // add the ID to the pool delete &e; // delete the entity } As you can see on the lines where I am telling the ComponentManager and the EntityManager that an Entity will die, I am calling a method to make sure it handles it appropriately. Now I realise I could do this without calling it explicitly, with the help of that for loop notifying all listener objects connected to the EntityFactory's Mouth (an object used to tell listeners that there's an event), however is this a good idea (good design, or what)? I've gone over the PROS and CONS, I just can't decide what I want to do. Calling Explicitly: PROS Faster? Since these functions are explicitly called, they can't be "removed" CONS Not flexible Bad design? (friend functions) Calling through Listener objects (i.e. ComponentManager/EntityManager inherits from a EntityFactoryListener) PROS More Flexible? Better Design? CONS Slower? (virtual functions) Listeners can be removed, i.e. may be removed and not get called again during the program, which could cause in a crash. P.S. If you wish to view my current source code, I am hosting it on BitBucket.

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• How to use correctly the comments in C/++

  - by Lucio

  I'm learning to program in C and in my stage, the best form to use correctly the comments is writing good comments from the beginning. As the comments are not just for that one understands better the code but others too, I want to know the views of all of you to reach a consensus. So what I want is that the most experienced users edit the following code as you please. (If it's unnecessary, delete it; If it's wrong, correct it; If needed, add more) Thus there'll be multiple answers with different syntax and the responses with the most votes will be taken as referring when commenting. The code to copy, paste and edit to your pleasure is: (And I remark again, just import the comments, not the code) /* This programs find 1 number in 1 file. The file is binary type and has integers in series. The number is integer type and it's entered from the keyboard. When finished the program, a poster will show the results: Saying if the number is in the file or not. */ #include <stdio.h> //FUNCTION 1 //Open file 'path' and closes it. void openf(char path[]) { int num; //Read from Keyboard a Number and it save it into 'num' var printf("Ready for read number.\n\nNumber --> "); fflush(stdin); scanf("%d",&num); //Open file 'path' in READ mode FILE *fvar; fvar=fopen(path,"rb"); //IF error happens when open file, exit of function if (fvar==NULL) { printf("ERROR while open file %s in read mode.",path); exit(1); } /*Verify the result of 'funct' function IF TRUE, 'num' it's in the file*/ if (funct(path,fvar,num)) printf("The number %d it is in the file %s.",num,path); else printf("The number %d it is not in the file %s.",num,path); fclose(fvar); } /*FUNCTION 2 It is a recursive function. Reads number by number until the file is empty or the number is found. Parameters received: 'path' -> Directory file 'fvar' -> Pointer file 'num' -> Number to compare */ int funct(char path[],FILE *fvar,int num) { int compare; //FALSE condition when the pointer reaches the end if (fread(&compare,sizeof(int),1,fvar)>0) /*TRUE condition when the number readed is iqual that 'num' ELSE will go to the function itself*/ if (compare!=num) funct(path,fvar,num); else return 1; else return 0; } int main(int argc, char **argv) { char path[30]="file.bin"; //Direction of the file to process openf(path); //Function with algorithm return 0; }

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• JDeveloper 11.1.2 : Command Link in Table Column Work Around

  - by Frank Nimphius

  Just figured that in Oracle JDeveloper 11.1.2, clicking on a command link in a table does not mark the table row as selected as it is the behavior in previous releases of Oracle JDeveloper. For the time being, the following work around can be used to achieve the "old" behavior: To mark the table row as selected, you need to build and queue the table selection event in the code executed by the command link action listener. To queue a selection event, you need to know about the rowKey of the row that the command link that you clicked on is located in. To get to this information, you add an f:attribute tag to the command link as shown below <af:column sortProperty="#{bindings.DepartmentsView1.hints.DepartmentId.name}" sortable="false"    headerText="#{bindings.DepartmentsView1.hints.DepartmentId.label}" id="c1">   <af:commandLink text="#{row.DepartmentId}" id="cl1" partialSubmit="true"       actionListener="#{BrowseBean.onCommandItemSelected}">     <f:attribute name="rowKey" value="#{row.rowKey}"/>   </af:commandLink>   ... </af:column> The f:attribute tag references #{row.rowKey} wich in ADF translates to JUCtrlHierNodeBinding.getRowKey(). This information can be used in the command link action listener to compose the RowKeySet you need to queue the selected row. For simplicitly reasons, I created a table "binding" reference to the managed bean that executes the command link action. The managed bean code that is referenced from the af:commandLink actionListener property is shown next: public void onCommandItemSelected(ActionEvent actionEvent) {   //get access to the clicked command link   RichCommandLink comp = (RichCommandLink)actionEvent.getComponent();   //read the added f:attribute value   Key rowKey = (Key) comp.getAttributes().get("rowKey");     //get the current selected RowKeySet from the table   RowKeySet oldSelection = table.getSelectedRowKeys();   //build an empty RowKeySet for the new selection   RowKeySetImpl newSelection = new RowKeySetImpl();     //RowKeySets contain List objects with key objects in them   ArrayList list = new ArrayList();   list.add(rowKey);   newSelection.add(list);     //create the selectionEvent and queue it   SelectionEvent selectionEvent = new SelectionEvent(oldSelection, newSelection, table);   selectionEvent.queue();     //refresh the table   AdfFacesContext.getCurrentInstance().addPartialTarget(table); }

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• Utility to Script SQL Server Configuration

  - by Bill Graziano

  I wrote a small utility to script some key SQL Server configuration information. I had two goals for this utility: Assist with disaster recovery preparation Identify configuration changes I’ve released the application as open source through CodePlex. You can download it from CodePlex at the Script SQL Server Configuration project page. The application is a .NET 2.0 console application that uses SMO. It writes its output to a directory that you specify.  Disaster Planning ScriptSqlConfig generates scripts for logins, jobs and linked servers.  It writes the properties and configuration from the instance to text files. The scripts are designed so they can be run against a DR server in the case of a disaster. The properties and configuration will need to be manually compared. Each job is scripted to its own file. Each linked server is scripted to its own file. The linked servers don’t include the password if you use a SQL Server account to connect to the linked server. You’ll need to store those somewhere secure. All the logins are scripted to a single file. This file includes windows logins, SQL Server logins and any server role membership.  The SQL Server logins are scripted with the correct SID and hashed passwords. This means that when you create the login it will automatically match up to the users in the database and have the correct password. This is the only script that I programmatically generate rather than using SMO. The SQL Server configuration and properties are scripted to text files. These will need to be manually reviewed in the event of a disaster. Or you could DIFF them with the configuration on the new server. Configuration Changes These scripts and files are all designed to be checked into a version control system.  The scripts themselves don’t include any date specific information. In my environments I run this every night and check in the changes. I call the application once for each server and script each server to its own directory.  The process will delete any existing files before writing new ones. This solved the problem I had where the scripts for deleted jobs and linked servers would continue to show up.  To see any changes I just need to query the version control system to show many any changes to the files. Database Scripting Utilities that script database objects are plentiful.  CodePlex has at least a dozen of them including one I wrote years ago. The code is so easy to write it’s hard not to include that functionality. This functionality wasn’t high on my list because it’s included in a database backup.  Unless you specify the /nodb option, the utility will script out many user database objects. It will script one object per file. It will script tables, stored procedures, user-defined data types, views, triggers, table types and user-defined functions. I know there are more I need to add but haven’t gotten around it yet. If there’s something you need, please log an issue and get it added. Since it scripts one object per file these really aren’t appropriate to recreate an empty database. They are really good for checking into source control every night and then seeing what changed. I know everyone tells me all their database objects are in source control but a little extra insurance never hurts. Conclusion I hope this utility will help a few of you out there. My goal is to have it script all server objects that aren’t contained in user databases. This should help with configuration changes and especially disaster recovery.

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• Grub2 : Windows 7 can't boot installing with Ubuntu 10.04 on different hard drive

  - by dellphi

  I use a dual boot with two hard disks and two OS is Ubuntu 10.04 and Windows 7. Windows 7 installed on the first disk, first partition. Grub is installed on a second hard disk MBR, and Ubuntu installed on an extended partition on a second hard drive. When I select Windows 7 on the Grub menu, the HDD lamp lights up briefly and then black screen on the monitor, with the status of the keyboard is still functioning. Until now (with the default boot from first HDD), I have to press F12 to get into the Grub to run Linux on a second HDD. ================ fdisk -l ================================ dellph1@dellph1-desktop:~$ fdisk -l omitting empty partition (5) Disk /dev/sda: 1000.2 GB, 1000204886016 bytes 255 heads, 63 sectors/track, 121601 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00087dec Device Boot Start End Blocks Id System /dev/sda1 * 1 23104 185582848+ 7 HPFS/NTFS /dev/sda2 23105 121601 791177122 5 Extended /dev/sda5 36107 74408 307660783+ 7 HPFS/NTFS /dev/sda6 74409 100081 206218341 7 HPFS/NTFS /dev/sda7 100082 121601 172859368+ 7 HPFS/NTFS Disk /dev/sdb: 160.0 GB, 160041885696 bytes 255 heads, 63 sectors/track, 19457 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x6d43dfb2 Device Boot Start End Blocks Id System /dev/sdb1 1 10030 80560066 5 Extended /dev/sdb5 * 1 5560 44657601 83 Linux /dev/sdb6 5560 9387 30736384 83 Linux /dev/sdb7 9387 10030 5164032 82 Linux swap / Solaris dellph1@dellph1-desktop:~$ ================= grub.cfg ================== # DO NOT EDIT THIS FILE # It is automatically generated by /usr/sbin/grub-mkconfig using templates from /etc/grub.d and settings from /etc/default/grub # BEGIN /etc/grub.d/00_header if [ -s $prefix/grubenv ]; then load_env fi set default="0" if [ ${prev_saved_entry} ]; then set saved_entry=${prev_saved_entry} save_env saved_entry set prev_saved_entry= save_env prev_saved_entry set boot_once=true fi function savedefault { if [ -z ${boot_once} ]; then saved_entry=${chosen} save_env saved_entry fi } function recordfail { set recordfail=1 if [ -n ${have_grubenv} ]; then if [ -z ${boot_once} ]; then save_env recordfail; fi; fi } insmod ext2 set root='(hd1,5)' search --no-floppy --fs-uuid --set 2f014a3a-35f3-4d05-87aa-34ca677160b7 if loadfont /usr/share/grub/unicode.pf2 ; then set gfxmode=1024x768 insmod gfxterm insmod vbe if terminal_output gfxterm ; then true ; else # For backward compatibility with versions of terminal.mod that don't # understand terminal_output terminal gfxterm fi fi insmod ext2 set root='(hd1,5)' search --no-floppy --fs-uuid --set 2f014a3a-35f3-4d05-87aa-34ca677160b7 set locale_dir=($root)/boot/grub/locale set lang=en insmod gettext if [ ${recordfail} = 1 ]; then set timeout=-1 else set timeout=5 fi END /etc/grub.d/00_header BEGIN /etc/grub.d/05_debian_theme insmod ext2 set root='(hd1,5)' search --no-floppy --fs-uuid --set 2f014a3a-35f3-4d05-87aa-34ca677160b7 insmod jpeg if background_image /usr/share/backgrounds/CurlsbyCandy.jpg ; then set color_normal=white/black set color_highlight=black/light-gray else set menu_color_normal=white/black set menu_color_highlight=black/light-gray fi END /etc/grub.d/05_debian_theme BEGIN /etc/grub.d/10_linux menuentry 'Ubuntu, with Linux 2.6.32-24-generic' --class ubuntu --class gnu-linux --class gnu --class os { recordfail insmod ext2 set root='(hd1,5)' search --no-floppy --fs-uuid --set 2f014a3a-35f3-4d05-87aa-34ca677160b7 linux /boot/vmlinuz-2.6.32-24-generic root=UUID=2f014a3a-35f3-4d05-87aa-34ca677160b7 ro splash vga=795 quiet splash nomodeset video=uvesafb:mode_option=1280x1024-24,mtrr=3,scroll=ywrap initrd /boot/initrd.img-2.6.32-24-generic } menuentry 'Ubuntu, with Linux 2.6.32-24-generic (recovery mode)' --class ubuntu --class gnu-linux --class gnu --class os { recordfail insmod ext2 set root='(hd1,5)' search --no-floppy --fs-uuid --set 2f014a3a-35f3-4d05-87aa-34ca677160b7 echo 'Loading Linux 2.6.32-24-generic ...' linux /boot/vmlinuz-2.6.32-24-generic root=UUID=2f014a3a-35f3-4d05-87aa-34ca677160b7 ro single splash vga=795 echo 'Loading initial ramdisk ...' initrd /boot/initrd.img-2.6.32-24-generic } END /etc/grub.d/10_linux BEGIN /etc/grub.d/30_os-prober menuentry "Windows 7 (loader) (on /dev/sda1)" { insmod ntfs set root='(hd0,1)' search --no-floppy --fs-uuid --set 5cac2139ac210f58 chainloader +1 } END /etc/grub.d/30_os-prober BEGIN /etc/grub.d/40_multisystem Ajout de MultiSystem MULTISYSTEM MENU menuentry "PLoP Boot Manager" { linux16 /boot/plpbt } menuentry "Smart Boot Manager" { search --set -f /boot/sbootmgr.dsk linux16 /boot/memdisk initrd16 /boot/sbootmgr.dsk } FIN MULTISYSTEM MENU END /etc/grub.d/40_multisystem ================================================ I want to keep the Grub on the second HDD. I have been using the Startup Manager, Boot Manager and Grub Customizer, and this problem still unsolved. The easiest thing that I can possibly do is to install Grub on first HDD, but I was curious and maybe someone can help.

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• ORE graphics using Remote Desktop Protocol

  - by Sherry LaMonica

  Oracle R Enterprise graphics are returned as raster, or bitmap graphics. Raster images consist of tiny squares of color information referred to as pixels that form points of color to create a complete image. Plots that contain raster images render quickly in R and create small, high-quality exported image files in a wide variety of formats. However, it is a known issue that the rendering of raster images can be problematic when creating graphics using a Remote Desktop connection. Raster images do not display in the windows device using Remote Desktop under the default settings. This happens because Remote Desktop restricts the number of colors when connecting to a Windows machine to 16 bits per pixel, and interpolating raster graphics requires many colors, at least 32 bits per pixel.. For example, this simple embedded R image plot will be returned in a raster-based format using a standalone Windows machine:  R> library(ORE) R> ore.connect(user="rquser", sid="orcl", host="localhost", password="rquser", all=TRUE)  R> ore.doEval(function() image(volcano, col=terrain.colors(30))) Here, we first load the ORE packages and connect to the database instance using database login credentials. The ore.doEval function executes the R code within the database embedded R engine and returns the image back to the client R session. Over a Remote Desktop connection under the default settings, this graph will appear blank due to the restricted number of colors. Users who encounter this issue have two options to display ORE graphics over Remote Desktop: either raise Remote Desktop's Color Depth or direct the plot output to an alternate device. Option #1: Raise Remote Desktop Color Depth setting In a Remote Desktop session, all environment variables, including display variables determining Color Depth, are determined by the RCP-Tcp connection settings. For example, users can reduce the Color Depth when connecting over a slow connection. The different settings are 15 bits, 16 bits, 24 bits, or 32 bits per pixel. To raise the Remote Desktop color depth: On the Windows server, launch Remote Desktop Session Host Configuration from the Accessories menu.Under Connections, right click on RDP-Tcp and select Properties.On the Client Settings tab either uncheck LimitMaximum Color Depth or set it to 32 bits per pixel. Click Apply, then OK, log out of the remote session and reconnect.After reconnecting, the Color Depth on the Display tab will be set to 32 bits per pixel.  Raster graphics will now display as expected. For ORE users, the increased color depth results in slightly reduced performance during plot creation, but the graph will be created instead of displaying an empty plot. Option #2: Direct plot output to alternate device Plotting to a non-windows device is a good option if it's not possible to increase Remote Desktop Color Depth, or if performance is degraded when creating the graph. Several device drivers are available for off-screen graphics in R, such as postscript, pdf, and png. On-screen devices include windows, X11 and Cairo. Here we output to the Cairo device to render an on-screen raster graphic.  The grid.raster function in the grid package is analogous to other grid graphical primitives - it draws a raster image within the current plot's grid.  R> options(device = "CairoWin") # use Cairo device for plotting during the session R> library(Cairo) # load Cairo, grid and png libraries  R> library(grid) R> library(png)  R> res <- ore.doEval(function()image(volcano,col=terrain.colors(30))) # create embedded R plot  R> img <- ore.pull(res, graphics = TRUE)$img[[1]] # extract image  R> grid.raster(as.raster(readPNG(img)), interpolate = FALSE) # generate raster graph R> dev.off() # turn off first device   By default, the interpolate argument to grid.raster is TRUE, which means that what is actually drawn by R is a linear interpolation of the pixels in the original image. Setting interpolate to FALSE uses a sample from the pixels in the original image.A list of graphics devices available in R can be found in the Devices help file from the grDevices package: R> help(Devices)

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• If the model is validating the data, shouldn't it throw exceptions on bad input?

  - by Carlos Campderrós

  Reading this SO question it seems that throwing exceptions for validating user input is frowned upon. But who should validate this data? In my applications, all validations are done in the business layer, because only the class itself really knows which values are valid for each one of its properties. If I were to copy the rules for validating a property to the controller, it is possible that the validation rules change and now there are two places where the modification should be made. Is my premise that validation should be done on the business layer wrong? What I do So my code usually ends up like this: <?php class Person { private $name; private $age; public function setName($n) { $n = trim($n); if (mb_strlen($n) == 0) { throw new ValidationException("Name cannot be empty"); } $this->name = $n; } public function setAge($a) { if (!is_int($a)) { if (!ctype_digit(trim($a))) { throw new ValidationException("Age $a is not valid"); } $a = (int)$a; } if ($a < 0 || $a > 150) { throw new ValidationException("Age $a is out of bounds"); } $this->age = $a; } // other getters, setters and methods } In the controller, I just pass the input data to the model, and catch thrown exceptions to show the error(s) to the user: <?php $person = new Person(); $errors = array(); // global try for all exceptions other than ValidationException try { // validation and process (if everything ok) try { $person->setAge($_POST['age']); } catch (ValidationException $e) { $errors['age'] = $e->getMessage(); } try { $person->setName($_POST['name']); } catch (ValidationException $e) { $errors['name'] = $e->getMessage(); } ... } catch (Exception $e) { // log the error, send 500 internal server error to the client // and finish the request } if (count($errors) == 0) { // process } else { showErrorsToUser($errors); } Is this a bad methodology? Alternate method Should maybe I create methods for isValidAge($a) that return true/false and then call them from the controller? <?php class Person { private $name; private $age; public function setName($n) { $n = trim($n); if ($this->isValidName($n)) { $this->name = $n; } else { throw new Exception("Invalid name"); } } public function setAge($a) { if ($this->isValidAge($a)) { $this->age = $a; } else { throw new Exception("Invalid age"); } } public function isValidName($n) { $n = trim($n); if (mb_strlen($n) == 0) { return false; } return true; } public function isValidAge($a) { if (!is_int($a)) { if (!ctype_digit(trim($a))) { return false; } $a = (int)$a; } if ($a < 0 || $a > 150) { return false; } return true; } // other getters, setters and methods } And the controller will be basically the same, just instead of try/catch there are now if/else: <?php $person = new Person(); $errors = array(); if ($person->isValidAge($age)) { $person->setAge($age); } catch (Exception $e) { $errors['age'] = "Invalid age"; } if ($person->isValidName($name)) { $person->setName($name); } catch (Exception $e) { $errors['name'] = "Invalid name"; } ... if (count($errors) == 0) { // process } else { showErrorsToUser($errors); } So, what should I do? I'm pretty happy with my original method, and my colleagues to whom I have showed it in general have liked it. Despite this, should I change to the alternate method? Or am I doing this terribly wrong and I should look for another way?

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