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• Creating a Reverse Proxy with URL Rewrite for IIS

  - by OWScott

  There are times when you need to reverse proxy through a server. The most common example is when you have an internal web server that isn’t exposed to the internet, and you have a public web server accessible to the internet. If you want to serve up traffic from the internal web server, you can do this through the public web server by creating a tunnel (aka reverse proxy). Essentially, you can front the internal web server with a friendly URL, even hiding custom ports. For example, consider an internal web server with a URL of http://10.10.0.50:8111. You can make that available through a public URL like http://tools.mysite.com/ as seen in the following image. The URL can be made public or it can be used for your internal staff and have it password protected and/or locked down by IP address. This is easy to do with URL Rewrite and IIS. You will also need Application Request Routing (ARR) installed even though for a simple reverse proxy you won’t use most of ARR’s functionality. If you don’t already have URL Rewrite and ARR installed you can do so easily with the Web Platform Installer. A lot can be said about reverse proxies and many different situations and ways to route the traffic and handle different URL patterns. However, my goal here is to get you up and going in the easiest way possible. Then you can dig in deeper after you get the base configuration in place. URL Rewrite makes a reverse proxy very easy to set up. Note that the URL Rewrite Add Rules template doesn’t include Reverse Proxy at the server level. That’s not to say that you can’t create a server-level reverse proxy, but the URL Rewrite rules template doesn’t help you with that. Getting Started First you must create a website on your public web server that has the public bindings that you need. Alternately, you can use an existing site and route using conditions for certain traffic. After you’ve created your site then open up URL Rewrite at the site level. Using the “Add Rule(s)…” template that is opened from the right-hand actions pane, create a new Reverse Proxy rule. If you receive a prompt (the first time) that the proxy functionality needs to be enabled, select OK. This is telling you that a proxy can route traffic outside of your web server, which happens to be our goal in this case. Be aware that reverse proxy rules can be dangerous if you open sites from inside you network to the world, so just be aware of what you’re doing and why. The next and final step of the template asks a few questions. The first textbox asks the name of the internal web server. In our example, it’s 10.10.0.50:8111. This can be any URL, including a subfolder like internal.mysite.com/blog. Don’t include the http or https here. The template assumes that it’s not entered. You can choose whether to perform SSL Offloading or not. If you leave this checked then all requests to the internal server will be over HTTP regardless of the original web request. This can help with performance and SSL bindings if all requests are within a trusted network. If the network path between the two web servers is not completely trusted and safe then uncheck this. Next, the template enables you to create an outbound rule. This is used to rewrite links in the page to look like your public domain name rather than the internal domain name. Outbound rules have a lot of CPU overhead because the entire web content needs to be parsed and updated. However, if you need it, then it’s well worth the extra CPU hit on the web server. If you check the “Rewrite the domain names of the links in HTTP responses” checkbox then the From textbox will be filled in with what you entered for the inbound rule. You can enter your friendly public URL for the outbound rule. This will essentially replace any reference to 10.10.0.50:8111 (or whatever you enter) with tools.mysite.com in all <a>, <form>, and <img> tags on your site. That’s it! Well, there is a lot more that you can do, this but will give you the base configuration. You can now visit www.mysite.com on your public web server and it will serve up the site from your internal web server. You should see two rules show up; one inbound and one outbound. You can edit these, add conditions, and tweak them further as needed. One common issue that can occur without outbound rules has to do with compression. If you run into errors with the new proxied site, try turning off compression to confirm if that’s the issue. Here’s a link with details on how to deal with compression and outbound rules. I hope this was helpful to get started and to see how easy it is to create a simple reverse proxy using URL Rewrite for IIS.

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• SOA Suite 11g Dynamic Payload Testing with soapUI Free Edition

  - by Greg Mally

  Overview Many web service developers use soapUI for various tests like: smoke test, unit test, and load testing because you can get a free edition that is fairly robust. However, if you need to venture into more complex testing that requires a dynamic payload, then the free edition doesn't necessarily make it easy. This feature does exist in soapUI, but for obvious reasons it is in the Pro version. In this blog I will show you how to use soapUI free edition for dynamic payloads in a simplified example. Hopefully this will open the doors for you to expand into more complex scenarios. The following assumes that you have a working knowledge of soapUI and will not go into concepts like setting up a project etc. For the basics, please review the documentation for soapUI: http://www.soapui.org/Getting-Started/. Additionally, we will be using asynchronous web services and you can review the setup for this in my blog: SOA Suite 11g Asynchronous Testing with soapUI. Features in soapUI Free Edition Relating to this Topic The soapUI test tool provides a very feature rich environment that can do many things provided you are willing to go beyond point and click. For this example, we will be leveraging just a couple features for our dynamic payload example: Test Case Properties Scripting with Groovy Basically, we will be using a property as a global variable and we will manipulate that property using a Groovy script. Setting Up Our Property Properties are available throughout soapUI and here is a snippet from the soapUI website defining the locations: Projects : for handling Project scope values, for example a subscription ID TestSuite : for handling TestSuite scoped values, can be seen as "arguments" to a TestSuite TestCases : for handling TestCase scoped values, can be seen as "arguments" to a TestCase Properties TestStep : for providing local values/state within a TestCase Local TestStep properties : several TestStep types maintain their own list of properties specific to their functionality : DataSource, DataSink, Run TestCase MockServices : for handling MockService scoped values/arguments MockResponses : for handling MockResponse scoped values Global Properties : for handling Global properties, optionally from an external source For our example, we will be defining a custom property in a TestCase called SimpleAsyncPayload. The property can be created in either the Custom Properties tab located at the bottom of the Navigator panel when the TestCase is selected in the Navigator or the Properties label in the TestCase editor: Navigator Panel TestCase Editor You will notice that I set a value of “0” for the custom property. For this simplified example, we will need to retrieve that value and manipulate it prior to making the web service request invocation. In order to accomplish this, we will need to get Groovy ;) Let's Get Groovy We will now add a new Groovy Script step to the TestCase called Manipulate Payload: TestCase Editor > Append Step > Groovy Script Once we have added the Groovy Script step to our TestCase, we can open the Groovy Script editor to add the code to: Get the current value of the property we created called SimpleAsyncPayload. Convert the value of the property to an integer. Increment the value. Store the incremented value back into the TestCase property called SimpleAsyncPayload. The script should look something like the following: Groovy Script Editor – Manipulate Payload At this point we can test the script to see if it is working by simply running the TestCase (left-click on the green triangle in the upper left-hand corner of the TestCase editor). To verify if it ran correctly, we can look at the value of the SimpleAsyncPayload property which should now be 1: TestCase Editor – Run Results All that is left to complete the TestCase is to append another step of type Test Request. The information required to append the request is a name and an operation to invoke. In this example we will use the default name and select the SimpleAsyncBPELProcessBingd -> process as the operation (any other information being requested, simply use the defaults unless you are calling an asynchronous operation then do not add any assertions). We are now in familiar ground with the Test Request editor. Depending upon the type of operation you are invoking (synchronous or asynchronous), please update the request with the necessary information (e.g., callback information for asynchronous operations). We will now tweak the Test Request payload to retrieve the value of the SimpleAsyncPayload property. The soapUI editor makes this very simple: right-click in the payload and navigate to the property (e.g., right-click > Get Data.. > TestCase: [Groovy TestCase] > Property [SimpleAsyncPayload]): Test Request Editor – Insert Property Value Your payload should now look something like the following: Test Request Editor – Inserted Property Value Just like before, we are now ready to run the TestCase. If everything goes as expected we should see a response like the following: Message Viewer – Results of TestCase Run We are now setup to be able to run a stress test where the payload will change for each request. This simple example can be expanded to include multiple payload values, complex calculations in the scripts, or whatever can be done via the soapUI scripting. Hopefully you have found this useful and happy testing to you :)

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• Configure IPv6 on your Linux system (Ubuntu)

  After the presentation on IPv6 at the first event of the Emtel Knowledge Series and some recent discussion on social media networks with other geeks and Linux interested IT people here in Mauritius, I thought that I should give it a try (finally) and tweak my local network infrastructure. Honestly, I have been to busy with contractual project work and it never really occurred to me to set up IPv6 in my LAN. Well, the following paragraphs are going to shed some light on those aspects of modern computer and network technology. This is the first article in a series on IPv6 configuration: Configure IPv6 on your Linux system DHCPv6: Provide IPv6 information in your local network Enabling DNS for IPv6 infrastructure Accessing your web server via IPv6 Piece of advice: This is based on my findings on the internet while reading other people's helpful articles and going through a couple of man-pages on my local system. Let's embrace IPv6 The basic configuration on Linux is actually very simple as the kernel, operating system, and user-space programs support that protocol natively. If your system is ready to go for IP (aka: IPv4), then you are good to go for anything else. At least, I didn't have to install any additional packages on my system(s). We are going to assign a static IPv6 address to the system. Hence, we have to modify the definition of interfaces and check whether we have an inet6 entry specified. Open your favourite text editor and check the following entries (it should be at least similar to this): $ sudo nano /etc/network/interfaces auto eth0# IPv4 configurationiface eth0 inet static  address 192.168.1.2  network 192.168.1.0  netmask 255.255.255.0  broadcast 192.168.1.255# IPv6 configurationiface eth0 inet6 static  pre-up modprobe ipv6  address 2001:db8:bad:a55::2  netmask 64 Of course, you might have to adjust your interface device (eth0) or you might be interested to have multiple directives for additional devices (eth1, eth2, etc.). The auto instruction takes care that your device is enabled and configured during the booting phase. The use of the pre-up directive depends on your kernel configuration but in most scenarios this might be an optional line. Anyways, it doesn't hurt to have it enabled after all - just to be on the safe side. Next, either restart your network subsystem like so: $ sudo service networking restart Or you might prefer to do it manually with identical parameters, like so: $ sudo ifconfig eth0 inet6 add 2001:db8:bad:a55::2/64 In case that you're logged in remotely into your PC (ie. via ssh), it is highly advised to opt for the second choice and add the device manually. You can check your configuration afterwards with one of the following commands (depends on whether it is installed): $ sudo ifconfig eth0eth0      Link encap:Ethernet  HWaddr 00:21:5a:50:d7:94            inet addr:192.168.160.2  Bcast:192.168.160.255  Mask:255.255.255.0          inet6 addr: fe80::221:5aff:fe50:d794/64 Scope:Link          inet6 addr: 2001:db8:bad:a55::2/64 Scope:Global          UP BROADCAST RUNNING MULTICAST  MTU:1500  Metric:1 $ sudo ip -6 address show eth03: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qlen 1000    inet6 2001:db8:bad:a55::2/64 scope global        valid_lft forever preferred_lft forever    inet6 fe80::221:5aff:fe50:d794/64 scope link        valid_lft forever preferred_lft forever In both cases, it confirms that our network device has been assigned a valid IPv6 address. That's it in general for your setup on one system. But of course, you might be interested to enable more services for IPv6, especially if you're already running a couple of them in your IP network. More details are available on the official Ubuntu Wiki. Continue to configure your network to provide IPv6 address information automatically in your local infrastructure.

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• Beginner Guide to User Styles for Firefox

  - by Asian Angel

  While the default styles for most websites are nice there may be times when you would love to tweak how things look. See how easy it can be to change how websites look with the Stylish Extension for Firefox. Note: Scripts from Userstyles.org can also be added to Greasemonkey if you have it installed. Getting Started After installing the extension you will be presented with a first run page. You may want to keep it open so that you can browse directly to the Userstyles.org website using the link in the upper left corner. In the lower right corner you will have a new Status Bar Icon. If you have used Greasemonkey before this icon works a little differently. It will be faded out due to no user style scripts being active at the moment. You can use either a left or right click to access the Context Menu. The user style script management section is also added into your Add-ons Management Window instead of being separate. When you reach the user style scripts homepage you can choose to either learn more about the extension & scripts or… Start hunting for lots of user style script goodness. There will be three convenient categories to get you jump-started if you wish. You could also conduct a search if you have something specific in mind. Here is some information directly from the website provided for your benefit. Notice the reference to using these scripts with Greasemonkey… This section shows you how the scripts have been categorized and can give you a better idea of how to search for something more specific. Finding & Installing Scripts For our example we decided to look at the Updated Styles Section”first. Based on the page number listing at the bottom there are a lot of scripts available to look through. Time to refine our search a little bit… Using the drop-down menu we selected site styles and entered Yahoo in the search blank. Needless to say 5 pages was a lot easier to look through than 828. We decided to install the Yahoo! Result Number Script. When you do find a script (or scripts) that you like simply click on the Install with Stylish Button. A small window will pop up giving you the opportunity to preview, proceed with the installation, edit the code, or cancel the process. Note: In our example the Preview Function did not work but it may be something particular to the script or our browser’s settings. If you decide to do some quick editing the window shown above will switch over to this one. To return to the previous window and install the user style script click on the Switch to Install Button. After installing the user style the green section in the script’s webpage will actually change to this message… Opening up the Add-ons Manager Window shows our new script ready to go. The script worked perfectly when we conducted a search at Yahoo…the Status Bar Icon also changed from faded out to full color (another indicator that everything is running nicely). Conclusion If you prefer a custom look for your favorite websites then you can have a lot of fun experimenting with different user style scripts. Note: See our article here for specialized How-To Geek User Style Scripts that can be added to your browser. Links Download the Stylish Extension (Mozilla Add-ons) Visit the Userstyles.org Website Install the Yahoo! Result Number User Style Similar Articles Productive Geek Tips Spice Up that Boring about:blank Page in FirefoxExpand the Add Bookmark Dialog in Firefox by DefaultEnjoy How-To Geek User Style Script GoodnessAuto-Hide Your Cluttered Firefox Status Bar ItemsBeginner Geek: Delete User Accounts in Windows 7 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 VMware Workstation 7 Acronis Online Backup DVDFab 6 Revo Uninstaller Pro Bypass Waiting Time On Customer Service Calls With Lucyphone MELTUP – "The Beginning Of US Currency Crisis And Hyperinflation" Enable or Disable the Task Manager Using TaskMgrED Explorer++ is a Worthy Windows Explorer Alternative Error Goblin Explains Windows Error Codes Twelve must-have Google Chrome plugins

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• Subterranean IL: Generics and array covariance

  - by Simon Cooper

  Arrays in .NET are curious beasts. They are the only built-in collection types in the CLR, and SZ-arrays (single dimension, zero-indexed) have their own commands and IL syntax. One of their stranger properties is they have a kind of built-in covariance long before generic variance was added in .NET 4. However, this causes a subtle but important problem with generics. First of all, we need to briefly recap on array covariance. SZ-array covariance To demonstrate, I'll tweak the classes I introduced in my previous posts: public class IncrementableClass { public int Value; public virtual void Increment(int incrementBy) { Value += incrementBy; } } public class IncrementableClassx2 : IncrementableClass { public override void Increment(int incrementBy) { base.Increment(incrementBy); base.Increment(incrementBy); } } In the CLR, SZ-arrays of reference types are implicitly convertible to arrays of the element's supertypes, all the way up to object (note that this does not apply to value types). That is, an instance of IncrementableClassx2[] can be used wherever a IncrementableClass[] or object[] is required. When an SZ-array could be used in this fashion, a run-time type check is performed when you try to insert an object into the array to make sure you're not trying to insert an instance of IncrementableClass into an IncrementableClassx2[]. This check means that the following code will compile fine but will fail at run-time: IncrementableClass[] array = new IncrementableClassx2[1]; array[0] = new IncrementableClass(); // throws ArrayTypeMismatchException These checks are enforced by the various stelem* and ldelem* il instructions in such a way as to ensure you can't insert a IncrementableClass into a IncrementableClassx2[]. For the rest of this post, however, I'm going to concentrate on the ldelema instruction. ldelema This instruction pops the array index (int32) and array reference (O) off the stack, and pushes a pointer (&) to the corresponding array element. However, unlike the ldelem instruction, the instruction's type argument must match the run-time array type exactly. This is because, once you've got a managed pointer, you can use that pointer to both load and store values in that array element using the ldind* and stind* (load/store indirect) instructions. As the same pointer can be used for both input and output to the array, the type argument to ldelema must be invariant. At the time, this was a perfectly reasonable restriction, and maintained array type-safety within managed code. However, along came generics, and with it the constrained callvirt instruction. So, what happens when we combine array covariance and constrained callvirt? .method public static void CallIncrementArrayValue() { // IncrementableClassx2[] arr = new IncrementableClassx2[1] ldc.i4.1 newarr IncrementableClassx2 // arr[0] = new IncrementableClassx2(); dup newobj instance void IncrementableClassx2::.ctor() ldc.i4.0 stelem.ref // IncrementArrayValue<IncrementableClass>(arr, 0) // here, we're treating an IncrementableClassx2[] as IncrementableClass[] dup ldc.i4.0 call void IncrementArrayValue<class IncrementableClass>(!!0[],int32) // ... ret } .method public static void IncrementArrayValue<(IncrementableClass) T>( !!T[] arr, int32 index) { // arr[index].Increment(1) ldarg.0 ldarg.1 ldelema !!T ldc.i4.1 constrained. !!T callvirt instance void IIncrementable::Increment(int32) ret } And the result: Unhandled Exception: System.ArrayTypeMismatchException: Attempted to access an element as a type incompatible with the array. at IncrementArrayValue[T](T[] arr, Int32 index) at CallIncrementArrayValue() Hmm. We're instantiating the generic method as IncrementArrayValue<IncrementableClass>, but passing in an IncrementableClassx2[], hence the ldelema instruction is failing as it's expecting an IncrementableClass[]. On features and feature conflicts What we've got here is a conflict between existing behaviour (ldelema ensuring type safety on covariant arrays) and new behaviour (managed pointers to object references used for every constrained callvirt on generic type instances). And, although this is an edge case, there is no general workaround. The generic method could be hidden behind several layers of assemblies, wrappers and interfaces that make it a requirement to use array covariance when calling the generic method. Furthermore, this will only fail at runtime, whereas compile-time safety is what generics were designed for! The solution is the readonly. prefix instruction. This modifies the ldelema instruction to ignore the exact type check for arrays of reference types, and so it lets us take the address of array elements using a covariant type to the actual run-time type of the array: .method public static void IncrementArrayValue<(IncrementableClass) T>( !!T[] arr, int32 index) { // arr[index].Increment(1) ldarg.0 ldarg.1 readonly. ldelema !!T ldc.i4.1 constrained. !!T callvirt instance void IIncrementable::Increment(int32) ret } But what about type safety? In return for ignoring the type check, the resulting controlled mutability pointer can only be used in the following situations: As the object parameter to ldfld, ldflda, stfld, call and constrained callvirt instructions As the pointer parameter to ldobj or ldind* As the source parameter to cpobj In other words, the only operations allowed are those that read from the pointer; stind* and similar that alter the pointer itself are banned. This ensures that the array element we're pointing to won't be changed to anything untoward, and so type safety within the array is maintained. This is a typical example of the maxim that whenever you add a feature to a program, you have to consider how that feature interacts with every single one of the existing features. Although an edge case, the readonly. prefix instruction ensures that generics and array covariance work together and that compile-time type safety is maintained. Tune in next time for a look at the .ctor generic type constraint, and what it means.

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• Disaster Recovery Discovery

  - by Rodney Landrum

  Last weekend I joined several of my IT staff on a mission to perform a DR test in our remote CoLo center in a large South East city of the US. Can I be more obtuse? The goal was simple for me as the sole DBA in a throng of Windows, Storage, Network and SAN admins – restore the databases and make them work. There were 4 applications that back ended to 7 SQL Server databases on 4 different SQL Server instances. We would maintain the original server names, but beyond that it was fair game. We had time to prepare so I was able to script out or otherwise automate the recovery process. I used sp_help_revlogin for three of the servers, a bit of a cheat actually because restoring the Master database on the target DR servers was the specified course of action according to the DR procedures ( the caveat “IF REQUIRED” left it open to interpretation. I really wanted to avoid the step of restoring Master for a number of reasons but mainly because I did not want to deal with issues starting SQL Services afterward. Having to account for the location of TempDB and the version conflicts of the resource DBs were just two of the battles I chose not to fight. Not to mention other system database location problems that might arise and prevent SQL from starting.  I was going to have to restore all of the user databases anyway, so I would not really gain any benefit, outside of logins, for taking the time to restore the source Master database over the newly installed one on the fresh server. What I wanted was the ability to restore the Master database as a user database, call it Master_Mine, from a backup on the source system and then use that restored database to script the SQL Logins and passwords on the DR systems. While I did not attempt this on the trip, the thought stuck in my mind and this past week I succeeded at scripting user accounts and passwords using only a restored copy of the Master database. Granted there were several challenges to overcome.  Also, as is usual for any work like this the usual disclaimers apply:  This is not something that I would imagine Microsoft would condone or support and this was really only an experiment for me to learn if it was even possible. While I have tested the process with success, I do not know that I would use this technique in a documented procedure because future updates for SQL Server will render this technique non-functional. I thought at first, incorrectly of course, that I could use sp_help_revlogin on a restored copy of the master database I named Master_Mine.   Since sp_help_revlogin uses system schema objects, sys.syslogins and sys.server_principals, this was not going to work because all results would come from the main Master database. To test this I added a SQL login via SSMS, backed up Master, restored  it as Master_Mine, and then deleted the login.  Even though the test account I created should presumably still be in the Master_Mine database, I should be able to get to it and script out its creation with its password hash so that I would not need to know the password, but any applications that stored that password would not have to be altered in the DR scenario. They would just work as expected. Once I realized that would not work I began looking deeper.  Knowing that sys.syslogins and sys.server_principals are system views, their underlying code should be available with sp_helptext, right? They were. And this led me to discover the two tables sys.sysxlgns and sys.sysprivs, where the data I needed was stored. These tables existed in both the real Master and the restored copy, Master_Mine.  I used this information to tweak the sp_help_revlogin stored procedure to use these tables instead to create the logins cursor used in sp_help_revlogin. For the password hash,  sp_help_revlogin uses the function LoginProperty() which takes a user name and option ‘passwordhash’ to return the hash for the user. Unfortunately, it requires the login to exist in the Master database. This would not work. So another slight modification I had to make was to pull the password hash itself (pwdhash from sys.sysxlgns) into the logins cursor and comment out the section of sp_help_revlogin that uses LoginProperty. Instead, I pass the pwdhash value as the variable @PWD_varbinary to the sp_hexadecimal stored procedure which is also created by and used within the code provided by Microsoft in the link above for sp_help_revlogin. The final challenge: sys.sysxlgns and sys.server_principals are visible only within a Dedicated Administrator Connection (DAC) query window in SSMS or within SQLCDMD.  To open a DAC connection you have to be logged in on the SQL Server itself, via RDP in my case,  and you preface the server name in the query connection with ADMIN:, so that the server connection looks like ADMIN:ServerName. From there you can create the modified stored procedure in the restored copy of a Master database from a source system as whatever name you like, and then run the modified stored procedure. I named my new stored procedure usp_help_revlogin_MyMaster. Upon execution I was happy to see the logins and password hashes that I needed to apply from the source Master database without having to restore over the new Master system database and without the need to access the original server (assuming it was down due to whatever disaster put it in that state). You will note that I am not providing full code samples here of the modifications. I will say that it was a slight bit of work and anyone who needed to do this for whatever reason, could fairly easily roll their own solution with the information provided herein.  My goal, as I said was to prove that this could be done and provide another option if required to ease the burden of getting SQL Servers up and available in an emergency situation where alternatives may be more challenging or otherwise unavailable.  

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• Sending Outlook Invites

  - by Daniel Moth

  Sending an Outlook invite for a meeting (also referred to as S+ in Microsoft) is a simple thing to get right if you just run the quick mental check below, which is driven by visual cues in the Outlook UI. I know that some folks don’t do this often or are new to Outlook, so if you know one of those folks share this blog post with them and if they read nothing else ask them to read step 7. Add on the To line the folks that you want to be at the meeting. Indicate optional invitees. Click on the “To” button to bring up the dialog that lets you move folks to be Optional (you can also do this from the Scheduling Assistant). Set the Reminder according to the attendee that has to travel the most. 5 minutes is the minimum. Use the Response Options and uncheck the "Request Response" if your event is going ahead regardless of who can make it or not, i.e. if everyone is optional. Don’t force every recipient to make an extra click, instead make the extra click yourself - you are the organizer. Add a good subject Make the subject such that just by reading it folks know what the meeting is about. Examples, e.g. "Review…", "Finalize…", "XYZ sync up" If this is only between two people and what is commonly referred to as a one to one, the subject would be something like "MyName/YourName 1:1" Write the subject in such a way that when the recipient sees this on their calendar among all the other items, they know what this meeting is about without having to see location, recipients, or any other information about the invite. Add a location, typically a meeting room. If recipients are from different buildings, schedule it where the folks that are doing the other folks a favor live. Otherwise schedule it wherever the least amount of people will have to travel. If you send me an invite to come to your building, and there is more of us than you, you are silently sending me the message that you are doing me a favor so if you don’t want to do that, include a note of why this is in your building, e.g. "Sorry we are slammed with back to back meetings today so hope you can come over to our building". If this is in someone's office, the location would be something like "Moth's office (7/666)" where in parenthesis you see the office location. If some folks are remote in another building/country, or if you know you picked a time which wasn't free for everyone, add an Online option (click the Lync Meeting button). Add a date and time. This MUST be at a time that is showing on the recipients’ calendar as FREE or at worst TENTATIVE. You can check that on the Scheduling Assistant. The reality is that this is not always possible, so in that case you MUST say something about it in the Invite Body, e.g. "Sorry I can see X has a conflict, but I cannot find a better slot", or "With so many of us there are some conflicts and I cannot find a better slot so hope this works", or "Apologies but due to Y we must have this meeting at this time and I know there are some conflicts, hope you can make it anyway". When you do that, I better not be able to find a better slot myself for all of us, and of course when you do that you have implicitly designated the Busy folks as optional. Finally, the body of the invite. This has the agenda of the meeting and if applicable the courtesy apologies due to messing up steps 6 & 7. This should not be the introduction to the meeting, in other words the recipients should not be surprised when they see the invite and go to the body to read it. Notifying them of the meeting takes place via separate email where you explain the purpose and give them a heads up that you'll be sending an invite. That separate email is also your chance to attach documents, don’t do that as part of the invite. TIP: If you have sent mail about the meeting, you can then go to your sent folder to select the message and click the "Meeting" button (Ctrl+Alt+R). This will populate the body with the necessary background, auto select the mandatory and optional attendees as per the TO/CC line, and have a subject that may be good enough already (or you can tweak it). Long to write, but very quick to remember and enforce since most of it is common sense and the checklist is driven of the visual cues in the UI you use to send the invite. Comments about this post by Daniel Moth welcome at the original blog.

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• Loading XML file containing leading zeros with SSIS preserving the zeros

  - by Compudicted

  Visiting the MSDN SQL Server Integration Services Forum oftentimes I could see that people would pop up asking this question: “why I am not able to load an element from an XML file that contains zeros so the leading/trailing zeros would remain intact?”. I started to suspect that such a trivial and often-required operation perhaps is being misunderstood by the developer community. I would also like to add that the whole state of affairs surrounding the XML today is probably also going to be increasingly affected by a motion of people who dislike XML in general and many aspects of it as XSD and XSLT invoke a negative reaction at best. Nevertheless, XML is in wide use today and its importance as a bridge between diverse systems is ever increasing. Therefore, I deiced to write up an example of loading an arbitrary XML file that contains leading zeros in one of its elements using SSIS so the leading zeros would be preserved keeping in mind the goal on simplicity into a table in SQL Server database. To start off bring up your BIDS (running as admin) and add a new Data Flow Task (DFT). This DFT will serve as container to adding our XML processing elements (besides, the XML Source is not available anywhere else other than from within the DFT). Double-click your DFT and drag and drop the XMS Source component from the Tool Box’s Data Flow Sources. Now, let the fun begin! Being inspired by the upcoming Christmas I created a simple XML file with one set of data that contains an imaginary SSN number of Rudolph containing several leading zeros like 0000003. This file can be viewed here. To configure the XML Source of course it is quite intuitive to point it to our XML file, next what the XML source needs is either an embedded schema (XSD) or it can generate one for us. In lack of the one I opted to auto-generate it for me and I ended up with an XSD that looked like: <?xml version="1.0"?> <xs:schema attributeFormDefault="unqualified" elementFormDefault="qualified" xmlns:xs="http://www.w3.org/2001/XMLSchema"> <xs:element name="XMasEvent"> <xs:complexType> <xs:sequence> <xs:element minOccurs="0" name="CaseInfo"> <xs:complexType> <xs:sequence> <xs:element minOccurs="0" name="ID" type="xs:unsignedByte" /> <xs:element minOccurs="0" name="CreatedDate" type="xs:unsignedInt" /> <xs:element minOccurs="0" name="LastName" type="xs:string" /> <xs:element minOccurs="0" name="FirstName" type="xs:string" /> <xs:element minOccurs="0" name="SSN" type="xs:unsignedByte" /> <!-- Becomes string -- > <xs:element minOccurs="0" name="DOB" type="xs:unsignedInt" /> <xs:element minOccurs="0" name="Event" type="xs:string" /> <xs:element minOccurs="0" name="ClosedDate" /> </xs:sequence> </xs:complexType> </xs:element> </xs:sequence> </xs:complexType> </xs:element> </xs:schema> As an aside on the XML file: if your XML file does not contain the outer node (<XMasEvent>) then you may end up in a situation where you see just one field in the output. Now please note that the SSN element’s data type was chosen to be of unsignedByte (and this is for a reason). The reason is stemming from the fact all our figures in the element are digits, this is good, but this is not exactly what we need, because if we will attempt to load the data with this XSD then we are going to either get errors on the destination or most typically lose the leading zeros. So the next intuitive choice is to change the data type to string. Besides, if a SSIS package was already created based on this XSD and the data type change was done thereafter, one should re-set the metadata by right-clicking the XML Source and choosing “Advanced Editor” in which there is a refresh button at the bottom left which will do the trick. So far so good, we are ready to load our XML file, well actually yes, and no, in my experience typically some data conversion may be required. So depending on your data destination you may need to tweak the data types targeted. Let’s add a Data Conversion Task to our DFT. Your package should look like: To make the story short I only will cover the SSN field, so in my data source the target SQL Table has it as nchar(10) and we chose string in our XSD (yes, this is a big difference), under such circumstances the SSIS will complain. So will go and manipulate on the data type of SSN by making it Unicode String (DT_WSTR), World String per se. The conversion should look like: The peek at the Metadata: We are almost there, now all we need is to configure the destination. For simplicity I chose SQL Server Destination. The mapping is a breeze, F5 and I am able to insert my data into SQL Server now! Checking the zeros – they are all intact!

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• Stumbling Through: Visual Studio 2010 (Part II)

  I would now like to expand a little on what I stumbled through in part I of my Visual Studio 2010 post and touch on a few other features of VS 2010.  Specifically, I want to generate some code based off of an Entity Framework model and tie it up to an actual data source.  Im not going to take the easy way and tie to a SQL Server data source, though, I will tie it to an XML data file instead.  Why?  Well, why not?  This is purely for learning, there are probably much better ways to get strongly-typed classes around XML but it will force us to go down a path less travelled and maybe learn a few things along the way.  Once we get this XML data and the means to interact with it, I will revisit data binding to this data in a WPF form and see if I cant get reading, adding, deleting, and updating working smoothly with minimal code.  To begin, I will use what was learned in the first part of this blog topic and draw out a data model for the MFL (My Football League) - I dont want the NFL to come down and sue me for using their name in this totally football-related article.  The data model looks as follows, with Teams having Players, and Players having a position and statistics for each season they played: Note that when making the associations between these entities, I was given the option to create the foreign key but I only chose to select this option for the association between Player and Position.  The reason for this is that I am picturing the XML that will contain this data to look somewhat like this: <MFL> <Position/> <Position/> <Position/> <Team>     <Player>         <Statistic/>     </Player> </Team> </MFL> Statistic will be under its associated Player node, and Player will be under its associated Team node no need to have an Id to reference it if we know it will always fall under its parent.  Position, however, is more of a lookup value that will not have any hierarchical relationship to the player.  In fact, the Position data itself may be in a completely different xml file (something Id like to play around with), so in any case, a player will need to reference the position by its Id. So now that we have a simple data model laid out, I would like to generate two things based on it:  A class for each entity with properties corresponding to each entity property An IO class with methods to get data for each entity, either all instances, by Id or by parent. Now my experience with code generation in the past has consisted of writing up little apps that use the code dom directly to regenerate code on demand (or using tools like CodeSmith).  Surely, there has got to be a more fun way to do this given that we are using the Entity Framework which already has built-in code generation for SQL Server support.  Lets start with that built-in stuff to give us a base to work off of.  Right click anywhere in the canvas of our model and select Add Code Generation Item: So just adding that code item seemed to do quite a bit towards what I was intending: It apparently generated a class for each entity, but also a whole ton more.  I mean a TON more.  Way too much complicated code was generated now that code is likely to be a black box anyway so it shouldnt matter, but we need to understand how to make this work the way we want it to work, so lets get ready to do some stumbling through that text template (tt) file. When I open the .tt file that was generated, right off the bat I realize there is going to be trouble there is no color coding, no intellisense no nothing!  That is going to make stumbling through more like groping blindly in the dark while handcuffed and hopping on one foot, which was one of the alternate titles I was considering for this blog.  Thankfully, the community comes to my rescue and I wont have to cast my mind back to the glory days of coding in VI (look it up, kids).  Using the Extension Manager (Available under the Tools menu), I did a quick search for tt editor in the Online Gallery and quickly found the Tangible T4 Editor: Downloading and installing this was a breeze, and after doing so I got some color coding and intellisense while editing the tt files.  If you will be doing any customizing of tt files, I highly recommend installing this extension.  Next, well see if that is enough help for us to tweak that tt file to do the kind of code generation that we wantDid you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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• Updates broke my themes/shell [Ubuntu 12.04 running Gnome 3 ]

  - by APNW

  I am running gnome-session 3.4.2.1. After the latest updates (listed below) my theme regressed to what looks like tango - not sure. Am unable to change it using Gnome-tweak tool or the display settings. I am also unable to change the wallpaper. Here's what it looks like: Synaptic: Chromium and this is the wallpaper page even though I have selected the wallpaper, it actually does not change. This same problem occurred on my personal computer, and one other computer I have, all running the same software/config. The interesting thing is that while Gnome 3 and Unity are affected, Cinnamon is not. What I've done so far: purged and re-installed both gnome 3 and Unity- no change noted. So, how do I fix this? Thanks Here's the installation log: Start-Date: 2013-11-07 12:01:28 Upgrade: chromium-browser-l10n:i386 (28.0.1500.71-0ubuntu1.12.04.1, 30.0.1599.114-0ubuntu0.12.04.3), libswscale2:i386 (0.8.6-0ubuntu0.12.04.1, 0.8.8-0ubuntu0.12.04.1), chromium-codecs-ffmpeg:i386 (28.0.1500.71-0ubuntu1.12.04.1, 30.0.1599.114-0ubuntu0.12.04.3), chromium-browser:i386 (28.0.1500.71-0ubuntu1.12.04.1, 30.0.1599.114-0ubuntu0.12.04.3), libpostproc52:i386 (0.8.6-0ubuntu0.12.04.1, 0.8.8-0ubuntu0.12.04.1), libavcodec-extra-53:i386 (0.8.6ubuntu0.12.04.1, 0.8.8ubuntu0.12.04.1), libavformat53:i386 (0.8.6-0ubuntu0.12.04.1, 0.8.8-0ubuntu0.12.04.1), libavutil-extra-51:i386 (0.8.6ubuntu0.12.04.1, 0.8.8ubuntu0.12.04.1) End-Date: 2013-11-07 12:02:00 Start-Date: 2013-11-07 17:32:55 Commandline: aptdaemon role='role-commit-packages' sender=':1.136' Install: libmusicbrainz5-0:i386 (5.0.1-2~precise2), udisks2:i386 (1.98.0-1~precise1), libclutter-gst-1.0-0:i386 (1.5.4-0ubuntu2), libudisks2-0:i386 (1.98.0-1~precise1), cinnamon-session-common:i386 (2.0.4-20131105043005-precise), librhythmbox-core6:i386 (2.97-1ubuntu1~precise1), gcr:i386 (3.4.1-3~precise1), libcluttergesture-0.0.2-0:i386 (0.0.2.1-2ubuntu3), libmx-1.0-2:i386 (1.4.3-0ubuntu1), guile-2.0-libs:i386 (2.0.5+1-1), libclutter-imcontext-0.1-0:i386 (0.1.4-2build1), libnatpmp1:i386 (20110808-3ubuntu1) Upgrade: gnome-keyring:i386 (3.2.2-2ubuntu4.1, 3.4.1-4ubuntu1~precise1), cinnamon:i386 (2.0.6-20131026040307-precise, 2.0.10-20131105040309-precise), gir1.2-muffin-3.0:i386 (2.0.3-20131023003029-precise, 2.0.3-20131105003012-precise), gir1.2-totem-1.0:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), nemo:i386 (2.0.2-20131023010018-precise, 2.0.5-20131105010007-precise), aisleriot:i386 (3.2.3.2-0ubuntu1, 3.4.1-1~precise1), procps:i386 (3.2.8-11ubuntu6.2, 3.2.8-11ubuntu6.3), libcinnamon-desktop0:i386 (2.0.2-20131025011504-precise, 2.0.3-20131105011505-precise), libgck-1-0:i386 (3.2.2-2ubuntu4.1, 3.4.1-3~precise1), totem-plugins:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), cinnamon-desktop-data:i386 (2.0.2-20131025011504-precise, 2.0.3-20131105011505-precise), rhythmbox:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), libgcr-3-1:i386 (3.2.2-2ubuntu4.1, 3.4.1-3~precise1), seahorse:i386 (3.2.2-0ubuntu2.1, 3.4.1-2~precise1), muffin-common:i386 (2.0.3-20131023003029-precise, 2.0.3-20131105003012-precise), totem-common:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), libtotem0:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), rhythmbox-data:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), gir1.2-cinnamondesktop-3.0:i386 (2.0.2-20131025011504-precise, 2.0.3-20131105011505-precise), cinnamon-session:i386 (2.0.1-20131021043004-precise, 2.0.4-20131105043005-precise), rhythmbox-mozilla:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), rhythmbox-plugin-zeitgeist:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), libmuffin0:i386 (2.0.3-20131023003029-precise, 2.0.3-20131105003012-precise), cjs:i386 (2.0.0-20131021020602-precise, 2.0.0-20131105020703-precise), rhythmbox-plugin-cdrecorder:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), cinnamon-common:i386 (2.0.6-20131026040307-precise, 2.0.10-20131105040309-precise), gnome-disk-utility:i386 (3.0.2-2ubuntu7, 3.4.1-0ubuntu1~precise1), nemo-fileroller:i386 (2.0.0-20131021020004-precise, 2.0.0-20131105020003-precise), libnemo-extension1:i386 (2.0.2-20131023010018-precise, 2.0.5-20131105010007-precise), rhythmbox-plugins:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), gimp:i386 (2.8.6-0precise1~ppa, 2.8.8-0precise0~ppa), cinnamon-settings-daemon:i386 (2.0.5-20131026004504-precise, 2.0.6-20131105004505-precise), libgimp2.0:i386 (2.8.6-0precise1~ppa, 2.8.8-0precise0~ppa), gir1.2-rb-3.0:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), wpasupplicant:i386 (0.7.3-6ubuntu2.1, 0.7.3-6ubuntu2.2), libcjs0c:i386 (2.0.0-20131021020602-precise, 2.0.0-20131105020703-precise), nemo-data:i386 (2.0.2-20131023010018-precise, 2.0.5-20131105010007-precise), totem:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), gimp-data:i386 (2.8.6-0precise1~ppa, 2.8.8-0precise0~ppa), transmission-common:i386 (2.51-0ubuntu1.3, 2.73-0ubuntu1~precise1), cinnamon-translations:i386 (2.0.1-20131021040407-precise, 2.0.1-20131105040807-precise), totem-mozilla:i386 (3.0.1-0ubuntu21.1, 3.4.3-0ubuntu1~precise1), rhythmbox-plugin-magnatune:i386 (2.96-0ubuntu4.3, 2.97-1ubuntu1~precise1), transmission-gtk:i386 (2.51-0ubuntu1.3, 2.73-0ubuntu1~precise1) End-Date: 2013-11-07 17:34:40

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• What are the industry metrics for average spend on dev hardware and software? [on hold]

  - by RationalGeek

  I'm trying to budget for my dev shop and compare our budget items to industry expectations. I'm hoping to find some information on what percentage of a dev's salary is generally spent on tooling, both hardware and software. Where can I find such information? If instead there is a source that looks at raw dollars that is useful, too. I can extrapolate what I need from that. NOTE: Your anecdotal evidence from your own job will not be very helpful. I'm looking for industry average statistics from a credible source. EDIT: I'm reluctant to even keep this question going based on the passionate negative responses of commenters, but I do think this is valuable information (assuming anyone will care to answer) so let me make one attempt to clarify why I'm looking for this information, and then leave it at that. I'm not sure why understanding and validating my motives is a necessary step to providing the information, but apparently that is the case, so I will do my best. Firstly, let me respond to the idea that us "management types" shouldn't use these types of metrics to evaluate budgets. I agree in part. Ideally, you should spend whatever is necessary on developers in order to keep them fully happy and productive. And this is true of all employees. However, companies operate in a world of limited resources, and every dollar spent in one area means a dollar not spent in another. So it is not enough to simply say "I need to spend $10,000 per developer next year" without having some way to justify that position. One way to help justify it is to compare yourself against the industry. If it is the case that on average a software shops spends 5% (making up that number) of their total development budget (salaries being the large portion of the other 95%, for arguments sake), and I'm only spending 3%, it helps in the justification process. So, it is not my intent to use this information to limit what I spend on developers, but rather to arm myself with the necessary justification to spend what I need to spend on developers to give them the best tools I can. I have been a developer for many years and I understand the need for proper tooling. Next, let's examine the idea that even considering the relationship between a spend on developer salaries and developer tooling is ludicrous and should be banned from budgetary thinking. As Jimmy Hoffa put it in their comment, it's like saying "I'm going to spend no more than 10% of median employee salary on light bulbs and coffee from now on.". Well, yes, it is like saying that, and from a budgeting perspective, this is a useful way to look at things. If you know that, on average, an employee consumes X dollars of coffee a year, then you can project a coffee budget based on that. And you can compare it to an industry metric to understand where you fall: do you spend more on coffee than other companies or less? Why might this be? If you are a coffee supply manager, that seems like a useful thought process. The same seems to hold true for developers. Now, on to the idea that I need to compare "apples to apples" and only look at other shops that are in the same place geographically, the same business, the same application architecture, and the same development frameworks. I guess if I could find such a statistic that said "a shop that is exactly identical to yours spends X on developer tooling" it would be wonderful. But there is plenty of value in an average statistic. Here's an analogy: let's say you are working on a household budget and need to decide how much to spend on groceries. Is it enough to know that the average consumer spends 15% on groceries and therefore decide that you will budget exactly 15%? No. You have to tweak your budget based on your individual needs and situation. But the generalized statistic does help in this evaluation. You can know if your budget is grossly off from what others are doing, and this can help you figure out why this is. So, I will concede the point that it would be better to find statistics that align to my shop, though I think any statistics I could find would be useful for what I'm doing. In that light, let's say that my shop is mostly focused on ASP.NET web applications. That doesn't map perfectly to reality because large enterprises have very heterogenous IT environments. But if I was going to pick one technology that is our focus that would be it. But, if you were to point me at some statistics that are related to a Linux shop doing embedded Java applications, I would still find it useful as a point of comparison. SUMMARY: Let me try to rephrase my question. I'm trying to find industry metrics on how much dev shops spend on developer tooling, both hardware and software. I don't so much care whether it is expressed as a percentage of total budget or as X dollars per dev or as Y percentage of salary. Any metric would be useful. If there are metrics that are specific to ASP.NET dev shops in the Northeast US, all the better, but I would be happy to find anything.

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• Stumbling Through: Visual Studio 2010 (Part II)

  I would now like to expand a little on what I stumbled through in part I of my Visual Studio 2010 post and touch on a few other features of VS 2010.  Specifically, I want to generate some code based off of an Entity Framework model and tie it up to an actual data source.  Im not going to take the easy way and tie to a SQL Server data source, though, I will tie it to an XML data file instead.  Why?  Well, why not?  This is purely for learning, there are probably much better ways to get strongly-typed classes around XML but it will force us to go down a path less travelled and maybe learn a few things along the way.  Once we get this XML data and the means to interact with it, I will revisit data binding to this data in a WPF form and see if I cant get reading, adding, deleting, and updating working smoothly with minimal code.  To begin, I will use what was learned in the first part of this blog topic and draw out a data model for the MFL (My Football League) - I dont want the NFL to come down and sue me for using their name in this totally football-related article.  The data model looks as follows, with Teams having Players, and Players having a position and statistics for each season they played: Note that when making the associations between these entities, I was given the option to create the foreign key but I only chose to select this option for the association between Player and Position.  The reason for this is that I am picturing the XML that will contain this data to look somewhat like this: <MFL> <Position/> <Position/> <Position/> <Team>     <Player>         <Statistic/>     </Player> </Team> </MFL> Statistic will be under its associated Player node, and Player will be under its associated Team node no need to have an Id to reference it if we know it will always fall under its parent.  Position, however, is more of a lookup value that will not have any hierarchical relationship to the player.  In fact, the Position data itself may be in a completely different xml file (something Id like to play around with), so in any case, a player will need to reference the position by its Id. So now that we have a simple data model laid out, I would like to generate two things based on it:  A class for each entity with properties corresponding to each entity property An IO class with methods to get data for each entity, either all instances, by Id or by parent. Now my experience with code generation in the past has consisted of writing up little apps that use the code dom directly to regenerate code on demand (or using tools like CodeSmith).  Surely, there has got to be a more fun way to do this given that we are using the Entity Framework which already has built-in code generation for SQL Server support.  Lets start with that built-in stuff to give us a base to work off of.  Right click anywhere in the canvas of our model and select Add Code Generation Item: So just adding that code item seemed to do quite a bit towards what I was intending: It apparently generated a class for each entity, but also a whole ton more.  I mean a TON more.  Way too much complicated code was generated now that code is likely to be a black box anyway so it shouldnt matter, but we need to understand how to make this work the way we want it to work, so lets get ready to do some stumbling through that text template (tt) file. When I open the .tt file that was generated, right off the bat I realize there is going to be trouble there is no color coding, no intellisense no nothing!  That is going to make stumbling through more like groping blindly in the dark while handcuffed and hopping on one foot, which was one of the alternate titles I was considering for this blog.  Thankfully, the community comes to my rescue and I wont have to cast my mind back to the glory days of coding in VI (look it up, kids).  Using the Extension Manager (Available under the Tools menu), I did a quick search for tt editor in the Online Gallery and quickly found the Tangible T4 Editor: Downloading and installing this was a breeze, and after doing so I got some color coding and intellisense while editing the tt files.  If you will be doing any customizing of tt files, I highly recommend installing this extension.  Next, well see if that is enough help for us to tweak that tt file to do the kind of code generation that we wantDid you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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• How can I load style resources from a dynamically loaded Silverlight application (XAP)?

  - by Tom

  I've followed Tim Heuer's video for dynamically loading other XAP's (into a 'master' Silverlight application), as well as some other links to tweak the loading of resources and am stuck on the particular issue of loading style resources from within the dynamically loaded XAP (i.e. the contents of Assets\Styles.xaml). When I run the master/hosting applcation, it successfully streams the dynamic XAP and I can read the deployment info etc. and load the assembly parts. However, when I actuall try to create an instance of a form from the Dynamic XAP, it fails with Cannot find a Resource with the Name/Key LayoutRootGridStyle which is in it's Assets\Styles.xaml file (it works if I run it directly so I know it's OK). For some reason these don't show up as application resources - not sure if I've totally got the wrong end of the stick, or am just missing something? Code snippet below (apologies it's a bit messy - just trying to get it working first) ... '' # Here's the code that reads the dynamic XAP from the web server ... '' #... wCli = New WebClient AddHandler wCli.OpenReadCompleted, AddressOf OpenXAPCompleted wCli.OpenReadAsync(New Uri("MyTest.xap", UriKind.Relative)) '' #... '' #Here's the sub that's called when openread is completed '' #... Private Sub OpenXAPCompleted(ByVal sender As Object, ByVal e As System.Net.OpenReadCompletedEventArgs) Dim sManifest As String = New StreamReader(Application.GetResourceStream(New StreamResourceInfo(e.Result, Nothing), New Uri("AppManifest.xaml", UriKind.Relative)).Stream).ReadToEnd Dim deploymentRoot As XElement = XDocument.Parse(sManifest).Root Dim deploymentParts As List(Of XElement) = _ (From assemblyParts In deploymentRoot.Elements().Elements() Select assemblyParts).ToList() Dim oAssembly As Assembly = Nothing For Each xElement As XElement In deploymentParts Dim asmPart As AssemblyPart = New AssemblyPart() Dim source As String = xElement.Attribute("Source").Value Dim sInfo As StreamResourceInfo = Application.GetResourceStream(New StreamResourceInfo(e.Result, "application/binary"), New Uri(source, UriKind.Relative)) If source = "MyTest.dll" Then oAssembly = asmPart.Load(sInfo.Stream) Else asmPart.Load(sInfo.Stream) End If Next Dim t As Type() = oAssembly.GetTypes() Dim AppClass = (From parts In t Where parts.FullName.EndsWith(".App") Select parts).SingleOrDefault() Dim mykeys As Array If Not AppClass Is Nothing Then Dim a As Application = DirectCast(oAssembly.CreateInstance(AppClass.FullName), Application) For Each strKey As String In a.Resources.Keys If Not Application.Current.Resources.Contains(strKey) Then Application.Current.Resources.Add(strKey, a.Resources(strKey)) End If Next End If Dim objectType As Type = oAssembly.GetType("MyTest.MainPage") Dim ouiel = Activator.CreateInstance(objectType) Dim myData As UIElement = DirectCast(ouiel, UIElement) Me.splMain.Children.Add(myData) Me.splMain.UpdateLayout() End Sub '' #... '' # And here's the line that fails with "Cannot find a Resource with the Name/Key LayoutRootGridStyle" '' # ... System.Windows.Application.LoadComponent(Me, New System.Uri("/MyTest;component/MainPage.xaml", System.UriKind.Relative)) '' #... any thoughts?

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• What might cause the big overhead of making a HttpWebRequest call?

  - by Dimitri C.

  When I send/receive data using HttpWebRequest (on Silverlight, using the HTTP POST method) in small blocks, I measure the very small throughput of 500 bytes/s over a "localhost" connection. When sending the data in large blocks, I get 2 MB/s, which is some 5000 times faster. Does anyone know what could cause this incredibly big overhead? Update: I did the performance measurement on both Firefox 3.6 and Internet Explorer 7. Both showed similar results. Update: The Silverlight client-side code I use is essentially my own implementation of the WebClient class. The reason I wrote it is because I noticed the same performance problem with WebClient, and I thought that the HttpWebRequest would allow to tweak the performance issue. Regrettably, this did not work. The implementation is as follows: public class HttpCommChannel { public delegate void ResponseArrivedCallback(object requestContext, BinaryDataBuffer response); public HttpCommChannel(ResponseArrivedCallback responseArrivedCallback) { this.responseArrivedCallback = responseArrivedCallback; this.requestSentEvent = new ManualResetEvent(false); this.responseArrivedEvent = new ManualResetEvent(true); } public void MakeRequest(object requestContext, string url, BinaryDataBuffer requestPacket) { responseArrivedEvent.WaitOne(); responseArrivedEvent.Reset(); this.requestMsg = requestPacket; this.requestContext = requestContext; this.webRequest = WebRequest.Create(url) as HttpWebRequest; this.webRequest.AllowReadStreamBuffering = true; this.webRequest.ContentType = "text/plain"; this.webRequest.Method = "POST"; this.webRequest.BeginGetRequestStream(new AsyncCallback(this.GetRequestStreamCallback), null); this.requestSentEvent.WaitOne(); } void GetRequestStreamCallback(IAsyncResult asynchronousResult) { System.IO.Stream postStream = webRequest.EndGetRequestStream(asynchronousResult); postStream.Write(requestMsg.Data, 0, (int)requestMsg.Size); postStream.Close(); requestSentEvent.Set(); webRequest.BeginGetResponse(new AsyncCallback(this.GetResponseCallback), null); } void GetResponseCallback(IAsyncResult asynchronousResult) { HttpWebResponse response = (HttpWebResponse)webRequest.EndGetResponse(asynchronousResult); Stream streamResponse = response.GetResponseStream(); Dim.Ensure(streamResponse.CanRead); byte[] readData = new byte[streamResponse.Length]; Dim.Ensure(streamResponse.Read(readData, 0, (int)streamResponse.Length) == streamResponse.Length); streamResponse.Close(); response.Close(); webRequest = null; responseArrivedEvent.Set(); responseArrivedCallback(requestContext, new BinaryDataBuffer(readData)); } HttpWebRequest webRequest; ManualResetEvent requestSentEvent; BinaryDataBuffer requestMsg; object requestContext; ManualResetEvent responseArrivedEvent; ResponseArrivedCallback responseArrivedCallback; } I use this code to send data back and forth to an HTTP server.

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• timeIntervalSinceDate Accuracy

  - by mmccomb

  I've been working on a game with an engine that updates 20 times per seconds. I've got to point now where I want to start getting some performance figures and tweak the rendering and logic updates. In order to do so I started to add some timing code to my game loop, implemented as follows... NSDate* startTime = [NSDate date]; // Game update logic here.... // Also timing of smaller internal events NSDate* endTime = [NSDate date]; [endTime timeIntervalSinceDate:startTime]; I noticed however that when I timed blocks within the outer timing logic that the time they took to execute did not sum up to match the overall time taken. So I wrote a small unit test to demonstrate the problem in which I time the overall time taken to complete the test and then 10 smaller events, here it is... - (void)testThatSumOfTimingsMatchesOverallTiming { NSDate* startOfOverallTime = [NSDate date]; // Variable to hold summation of smaller timing events in the upcoming loop... float sumOfIndividualTimes = 0.0; NSTimeInterval times[10] = {0.0, 0.0, 0.0, 0.0, 0.0, 0.0, 0.0, 0.0, 0.0, 0.0}; for (int i = 0; i < 10; i++) { NSDate* startOfIndividualTime = [NSDate date]; // Kill some time... sleep(1); NSDate* endOfIndividualTime = [NSDate date]; times[i] = [endOfIndividualTime timeIntervalSinceDate:startOfIndividualTime]; sumOfIndividualTimes += times[i]; } NSDate* endOfOverallTime = [NSDate date]; NSTimeInterval overallTimeTaken = [endOfOverallTime timeIntervalSinceDate:startOfOverallTime]; NSLog(@"Sum of individual times: %fms", sumOfIndividualTimes); NSLog(@"Overall time: %fms", overallTimeTaken); STAssertFalse(TRUE, @""); } And here's the output... Sum of individual times: 10.001377ms Overall time: 10.016834ms Which illustrates my problem quite clearly. The overall time was 0.000012ms but the smaller events took only 0.000001ms. So what happened to the other 0.000011ms? Is there anything that looks particularly wrong with my code? Or is there an alternative timing mechanism I should use?

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• Opera Mobile, offline web app development, and memory

  - by Jake Krohn

  I'm developing a data collection app for use on a HP iPAQ 211. I'm doing it as an offline web app (go with what you know) using Opera Mobile 9.7 and Google Gears. Being it is an offline app, it is very dependent on Javascript for much of its behavior. I'm using the LocalServer, Database, and Geolocation components of Gears, as well as the JQuery core and a couple of plugins for form validation and other usability tweaks (no jQuery UI). I've tried to be conservative with my programming style and free up or close resources whenever possible, but Opera just slowly dies after about 10-20 minutes of use. The Javascript engine stops responding, pages only half-load, and eventually stop loading completely. I'm guessing it's a resource issue. Quitting and relaunching the browser solves the problem, but only temporarily. The iPAQ ships with 128 MB of RAM, about 85-87 MB of which is available immediately after a reset. With only Opera running, there still remains about 50 MB that is left unused. My questions are thus: Is it possible to get Opera to address this unused RAM? Are there configuration settings in Opera or in the Windows Registry itself that will help improve performance? I know where to tweak, but the descriptions of the opera:config variables that I've found are less than helpful. Is is laughable to ask about memory management and jQuery in the same sentence? If not, does anyone have any suggestions? Finally, are my plans too ambitious, given the platform I have to work with? I know that Gears and Windows Mobile 6 are on their way out, but they (theoretically) suffice for what I need to do. I could ditch them in favor of an iPhone/iPod Touch, Mobile Safari, and HTML5 but I'd like to try to make this work first. I didn't think that Opera was a dog when it comes to JS performance, but perhaps it's worse than I thought. That this motley collection of technologies works at all is a minor miracle, but it needs to be faster and more stable. I appreciate any suggestions.

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• How to maximize http.sys file upload performance

  - by anelson

  I'm building a tool that transfers very large streaming data sets (possibly on the order of terabytes in a single stream; routinely in the tens of gigabytes) from one server to another. The client portion of the tool will read blocks from the source disk, and send them over the network. The server side will read these blocks off the network and write them to a file on the server disk. Right now I'm trying to decide which transport to use. Options are raw TCP, and HTTP. I really, REALLY want to be able to use HTTP. The HttpListener (or WCF if I want to go that route) make it easy to plug in to the HTTP Server API (http.sys), and I can get things like authentication and SSL for free. The problem right now is performance. I wrote a simple test harness that sends 128K blocks of NULL bytes using the BeginWrite/EndWrite async I/O idiom, with async BeginRead/EndRead on the server side. I've modified this test harness so I can do this with either HTTP PUT operations via HttpWebRequest/HttpListener, or plain old socket writes using TcpClient/TcpListener. To rule out issues with network cards or network pathways, both the client and server are on one machine and communicate over localhost. On my 12-core Windows 2008 R2 test server, the TCP version of this test harness can push bytes at 450MB/s, with minimal CPU usage. On the same box, the HTTP version of the test harness runs between 130MB/s and 200MB/s depending upon how I tweak it. In both cases CPU usage is low, and the vast majority of what CPU usage there is is kernel time, so I'm pretty sure my usage of C# and the .NET runtime is not the bottleneck. The box has two 6-core Xeon X5650 processors, 24GB of single-ranked DDR3 RAM, and is used exclusively by me for my own performance testing. I already know about HTTP client tweaks like ServicePointManager.MaxServicePointIdleTime, ServicePointManager.DefaultConnectionLimit, ServicePointManager.Expect100Continue, and HttpWebRequest.AllowWriteStreamBuffering. Does anyone have any ideas for how I can get HTTP.sys performance beyond 200MB/s? Has anyone seen it perform this well on any environment?

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• Convert JSON flattened for forms back to an object

  - by George Jempty

  I am required (please therefore no nit-picking the requirement, I've already nit-picked it, and this is the req) to convert certain form fields that have "object nesting" embedded in the field names, back to the object(s) themselves. Below are some typical form field names: phones_0_patientPhoneTypeId phones_0_phone phones_1_patientPhoneTypeId phones_1_phone The form fields above were derived from an object such as the one toward the bottom (see "Data"), and that is the format of the object I need to reassemble. It can be assumed that any form field with a name that contains the underscore _ character needs to undergo this conversion. Also that the segment of the form field between underscores, if numeric, signifies a Javascript array, otherwise an object. I found it easy to devise a (somewhat naive) implementation for the "flattening" of the original object for use by the form, but am struggling going in the other direction; below the object/data below I'm pasting my current attempt. One problem (perhaps the only one?) with it is that it does not currently properly account for array indexes, but this might be tricky because the object will subsequently be encoded as JSON, which will not account for sparse arrays. So if "phones_1" exists, but "phones_0" does not, I would nevertheless like to ensure that a slot exists for phones[0] even if that value is null. Implementations that tweak what I have begun, or are entirely different, encouraged. If interested let me know if you'd like to see my code for the "flattening" part that is working. Thanks in advance Data: var obj = { phones: [{ "patientPhoneTypeId": 4, "phone": "8005551212" }, { "patientPhoneTypeId": 2, "phone": "8885551212" }]}; Code to date: var unflattened = {}; for (var prop in values) { if (prop.indexOf('_') > -1) { var lastUnderbarPos = prop.lastIndexOf('_'); var nestedProp = prop.substr(lastUnderbarPos + 1); var nesting = prop.substr(0, lastUnderbarPos).split("_"); var nestedRef, isArray, isObject; for (var i=0, n=nesting.length; i<n; i++) { if (i===0) { nestedRef = unflattened; } if (i < (n-1)) { // not last if (/^\d+$/.test(nesting[i+1])) { isArray = true; isObject = false; } else { isArray = true; isObject = false; } var currProp = nesting[i]; if (!nestedRef[currProp]) { if (isArray) { nestedRef[currProp] = []; } else if (isObject) { nestedRef[currProp] = {}; } } nestedRef = nestedRef[currProp]; } else { nestedRef[nestedProp] = values[prop]; } } }

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• Loop through multi-dimensional array and remove certain keys

  - by Webkungen

  Hi! I've got a nested tree structure which is based on the array below: Array ( [1] = Array ( [id] = 1 [parent] = 0 [name] = Startpage [uri] = 125 [basename] = index.php [child] = ) [23] = Array ( [id] = 23 [parent] = 0 [name] = Events [uri] = 0 [basename] = [child] = Array ( [24] = Array ( [id] = 24 [parent] = 23 [name] = Public news [uri] = 0 [basename] = [child] = Array ( [27] = Array ( [id] = 27 [parent] = 24 [name] = Add [uri] = 100 [basename] = news.public.add.php [child] = ) [28] = Array ( [id] = 28 [parent] = 24 [name] = Overview [uri] = 101 [basename] = news.public.overview.php [child] = ) ) ) [25] = Array ( [id] = 25 [parent] = 23 [name] = Private news [uri] = 0 [basename] = [child] = Array ( [29] = Array ( [id] = 29 [parent] = 25 [name] = Add [uri] = 67 [basename] = news.private.add.php [child] = ) [30] = Array ( [id] = 30 [parent] = 25 [name] = Overview [uri] = 68 [basename] = news.private.overview.php [child] = ) ) ) [26] = Array ( [id] = 26 [parent] = 23 [name] = Calendar [uri] = 0 [basename] = [child] = Array ( [31] = Array ( [id] = 31 [parent] = 26 [name] = Add [uri] = 69 [basename] = news.event.add.php [child] = ) [32] = Array ( [id] = 32 [parent] = 26 [name] = Overview [uri] = 70 [basename] = news.event.overview.php [child] = ) ) ) ) ) ) I'm looking for a function to loop (recursive?) through the array and remove some keys. I my system I can allow users to certain functions/pages and if I deny access to the whole "block" "Events", the array will look like this: Array ( [1] = Array ( [id] = 1 [parent] = 0 [name] = Startpage [uri] = 125 [basename] = index.php [child] = ) [23] = Array ( [id] = 23 [parent] = 0 [name] = Events [uri] = 0 [basename] = [child] = Array ( [24] = Array ( [id] = 24 [parent] = 23 [name] = Public news [uri] = 0 [basename] = [child] = ) [25] = Array ( [id] = 25 [parent] = 23 [name] = Private news [uri] = 0 [basename] = [child] = ) [26] = Array ( [id] = 26 [parent] = 23 [name] = Calendar [uri] = 0 [basename] = [child] = ) ) ) ) As you can see above, the whole "block" "Events" is useless right now, becuase there is no page associated with each option. So I need to find all "keys" where "basename" is null AND where child is not an array or where the array is empty and remove them. I found this function when searching the site: function searchAndDestroy(&$a, $key, $val){ foreach($a as $k = &$v){ if(is_array($v)){ $r = searchAndDestroy($v, $key, $val); if($r){ unset($a[$k]); } }elseif($key == $k && $val == $v){ return true; } } return false; } It can be used to remove a key any where in the array, but only based in one thing, for example remove all keys where "parent" equals "23". But I need to find and remove (unset) all keys where "basename" is null AND where child isn't an array or where the array is empty. Can anyone help me out and possibly tweak the function above? Thank you,

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• Release Process Improvements

  - by wallismark

  The process of creating a new build and releasing it to production is a critical step in the SDLC but it is often left as an afterthought and varies greatly from one company to the next. I'm hoping people will share improvements they have made to this process in their organisation so we can all takes steps to 'reduce the pain'. So the question is, specify one painful/time consuming part of your release process and what did you do to improve it? My example: at a previous employer all developers made database changes on one common development database. Then when it came to release time, we used Redgate's SQL Compare to generate a huge script from the differences between the Dev and QA databases. This works reasonably well but the problems with this approach are:- ALL changes in the Dev database are included, some of which may still be 'works in progress'. Sometimes developers made conflicting changes (that were not noticed until the release was in production) It was a time consuming and manual process to create and validate the script (by validate I mean, try to weed out issues like problem 1 and 2). When there were problems with the script (eg the order in which things were run such as creating a record which relies on a foreign key record which is in the script but not yet run) it took time to 'tweak' it so it ran smoothly. It's not an ideal scenario for Continuous Integration. So the solution was:- Enforce a policy of all changes to the database must be scripted. A naming convention was important for ensuring the correct running order of the scripts. Create/Use a tool to run the scripts at release time. Developers had their own copy of the database do develop against (so there was no more 'stepping on each others toes') The next release after we started this process was much faster with fewer problems, indeed the only problems found were due to people 'breaking the rules', eg not creating a script. Once the issues with releasing to QA were fixed, when it came time to release to production it was very smooth. We applied a few other changes (like introducing CI) but this was the most significant, overall we reduced release time from around 3 hours down to a max of 10-15 minutes.

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• What's the correct way to use Cakephp urls?

  - by Pichan

  Hello all, it's my first post here :) I'm having some difficulties with dealing with urls and parameters. I've gone through the router class api documentation over and over again and found nothing useful. First of all, I'd like to know if there is any 'universal' format in CakePHP(1.3) for handling urls. I'm currently handling all my urls as simple arrays(in the format that Router::url and $html-link accepts) and it's easy as long as I only need to pass them as arguments to cake's own methods. It usually gets tricky if I need something else. Mainly I'm having problems with converting string urls to the basic array format. Let's say I want to convert $arrayUrl to string and than again into url: $arrayUrl=array('controller'=>'SomeController','action'=>'someAction','someValue'); $url=Router::url($arrayUrl); //$url is now '/path/to/site/someController/someAction/someValue' $url=Router::normalize($url); //remove '/path/to/site' $parsed=Router::parse($url); /*$parsed is now Array( [controller] => someController [action] => someAction [named] => Array() [pass] => Array([0] => someValue) [plugin] => ) */ That seems an awful lot of code to do something as simple as to convert between 2 core formats. Also, note that $parsed is still not in the same as $arrayUrl. Of course I could tweak $parsed manually and actually I've done that a few times as a quick patch but I'd like to get to the bottom of this. I also noticed that when using prefix routing, $this-params in controller has the prefix embedded in the action(i.e. [action] = 'admin_edit') and the result of Router::parse() does not. Both of course have the prefix in it's own key. To summarize, how do I convert an url between any of these 3(or 4, if you include the prefix thing) mentioned formats the right way? Of course it would be easy to hack my way through this, but I'd still like to believe that cake is being developed by a bunch of people who have a lot more experience and insight than me, so I'm guessing there's a good reason for this "perceived misbehavior". I've tried to present my problem as good as I can, but due to my rusty english skills, I had to take a few detours :) I'll explain more if needed.

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• Delphi LoadLibrary Failing to find DLL other directory - any good options?

  - by Chris Thornton

  Two Delphi programs need to load foo.dll, which contains some code that injects a client-auth certificate into a SOAP request. foo.dll resides in c:\fooapp\foo.dll and is normally loaded by c:\fooapp\foo.exe. That works fine. The other program needs the same functionality, but it resides in c:\program files\unwantedstepchild\sadapp.exe. Both aps load the DLL with this code: FOOLib := LoadLibrary('foo.dll'); ... If FOOLib <> 0 then begin FOOProc := GetProcAddress(FOOLib , 'xInjectCert'); FOOProc(myHttpRequest, Data, CertName); end; It works great for foo.exe, as the dll is right there. sadapp.exe fails to load the library, so FOOLib is 0, and the rest never gets called. The sadapp.exe program therefore silently fails to inject the cert, and when we test against production, it the cert is missing, do the connection fails. Obviously, we should have fully-qualified the path to the DLL. Without going into a lot of details, there were aspects of the testing that masked this problem until recently, and now it's basically too late to fix in code, as that would require a full regression test, and there isn't time for that. Since we've painted ourselves into a corner, I need to know if there are any options that I've overlooked. While we can't change the code (for this release), we CAN tweak the installer. I've found that placing c:\fooapp into the path works. As does adding a second copy of foo.dll directly into c:\program files\unwantedstepchild. c:\fooapp\foo.exe will always be running while sadapp.exe is running, so I was hoping that Windows would find it that way, but apparently not. Is there a way to tell Windows that I really want that same DLL? Maybe a manifest or something? This is the sort of "magic bullet" that I'm looking for. I know I can: Modify the windows path, probably in the installer. That's ugly. Add a second copy of the DLL, directly into the unwantedstepchild folder. Also ugly Delay the project while we code and test a proper fix. Unacceptable. Other? Thanks for any guidance, especially with "Other". I understand that this issue is not necessarily specific to Delphi. Thanks!

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• Move options between multiple dropdown lists

  - by Martha

  We currently have a form with the standard multi-select functionality of "here are the available options, here are the selected options, here are some buttons to move stuff back and forth." However, the client now wants the ability to not just select certain items, but to also categorize them. For example, given a list of books, they want to not just select the ones they own, but also the ones they've read, the ones they would like to read, and the ones they've heard about. (All examples fictional.) Thankfully, a selected item can only be in one category at a time. I can find many examples of moving items between listboxes, but not a single one for moving items between multiple listboxes. To add to the complication, the form needs to have two sets of list+categories, e.g. a list of movies that need to be categorized in addition to the aforementioned books. An additional problem is that sorting between lists is all well and good in the javascript-enabled world, but I can't really think of a good fallback interface for, say, mobile browsers. Maybe a pseudo-listbox with radio buttons next to each item? The master list of items will in general be very long - over 100 items, certainly, possibly many more. Any given category will most likely contain one or two selected items, but the possibility exists for a category to have dozens of selected items, or zero selected items. As far as OS and stuff, the site is in classic asp (quit snickering!), the server-side code is VBScript, and so far we've avoided the various Javascript libraries by the simple expedient of almost never using client-side scripting. This one form for this one client is currently the big exception. Give 'em an inch and they want a mile... Oh, and I have to add: I suck at Javascript, or really at any C-descendant language. Curly braces give me hives. I'd really, really like something I can just copy & paste into my page, maybe tweak some variable names, and never look at it again. A girl can dream, can't she? :)

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• Create Duplicate Records on SELECT for Calendar Date Range

  - by peterallcdn

  Hey all, I've built a pretty shnazzy calendar system but there is one tweak that I need to make so that I'm completely happy with it. My calendar has three tables: calevents - The calendared event. caldates - The occurrences and date-range of each occurrence for each event. calcats - The categories that can be applied to an event. The short: For each calevent, there can be many caldates, one for each occurrence of calevent. So a calevent that repeats weekly and spans 3 days might have caldates like this: date_id date_eid date_start date_end 2 37 2010-06-21 2010-06-23 3 37 2010-06-28 2010-06-30 7 37 2010-07-05 2010-07-07 9 37 2010-07-12 2010-07-14 What I want to do, is when selecting all the caldates for a specified month such as 2010-06, to return not just the two records above, but instead a record for each date in the range of date_start and date_end for each caldate. So if I searched for 2010-06, I would get: date_id date_eid date_start date_end date_day 2 37 2010-06-21 2010-06-23 2010-06-21 2 37 2010-06-21 2010-06-23 2010-06-22 2 37 2010-06-21 2010-06-23 2010-06-23 3 37 2010-06-28 2010-06-30 2010-06-28 3 37 2010-06-28 2010-06-30 2010-06-29 3 37 2010-06-28 2010-06-30 2010-06-30 The Long: The reason I want to do this, is so when displaying a list of events(calevents) for a specified month, an occurrence(caldates) of that event will be displayed for EACH of the days it spans. I could do this with php by looping through each day of the current month and displaying a copy of each caldate if the month day falls between date_start and date_end. But doing it this way will prevent me from using record pagination if needed. For example, if for a specified month the following caldates were returned: date_id date_eid date_start date_end 2 37 2010-06-21 2010-06-27 94 53 2010-06-09 2010-07-08 Doing record pagination would see this as only 2 records("rows"). But looping through them with PHP would generate 29 "rows". So, I figure if I use mysql to create each row instead of PHP, I can achieve the same thing AND still be able to use pagination if a month has a lot of events/dates. As far as performance goes, I'm not sure which option is more efficient. Both would send the same amount of info to the browser, so it's really only the work required to generate the info that matters. My current query which fetches all the occurrences for a specified month, and to make things just a little more complicated... joins them with their event and category, looks like this: $sql_to_execute = " SELECT date_id, date_eid, date_start, date_end, event_id, event_title, event_category, event_private, event_location, SUBSTRING_INDEX(event_detailsstripped, ' ', 40) AS event_detailsstripped, event_time, event_starttime, event_endtime, event_active, cat_colour FROM ( caldates LEFT JOIN calevents ON caldates.date_eid = calevents.event_id ) LEFT JOIN calcats ON calevents.event_category = calcats.cat_id WHERE date_start <= '".mysql_real_escape_string($dbi_list_end_date)."' AND date_end >= '".mysql_real_escape_string($dbi_list_start_date)."' ".$dbi_category." ORDER BY date_start ASC "; Any help or advice would be greatly appreciated! Thanks, Peter

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• Comparing Nested object properties using C#

  - by Kumar

  I have a method which compares two objects and returns a list of all the property names which are different. public static IList<string> GetDifferingProperties(object source, object) { var sourceType = source.GetType(); var sourceProperties = sourceType.GetProperties(); var targetType = target.GetType(); var targetProperties = targetType.GetProperties(); var properties = (from s in sourceProperties from t in targetProperties where s.Name == t.Name && s.PropertyType == t.PropertyType && s.GetValue(source,null) != t.GetValue(target,null) select s.Name).ToList(); return properties; } For example if I have two classes as follows: public class Address { public string AddressLine1 { get; set; } public string AddressLine2 { get; set; } public string City { get; set; } public string State { get; set; } public string Zip { get; set; } } public class Employee { public string FirstName { get; set; } public string MiddleName { get; set; } public string LastName { get; set; } public Address EmployeeAddress { get; set; } } I am trying to compare the following two employee instances: var emp1Address = new Address(); emp1Address.AddressLine1 = "Microsoft Corporation"; emp1Address.AddressLine2 = "One Microsoft Way"; emp1Address.City = "Redmond"; emp1Address.State = "WA"; emp1Address.Zip = "98052-6399"; var emp1 = new Employee(); emp1.FirstName = "Bill"; emp1.LastName = "Gates"; emp1.EmployeeAddress = emp1Address; var emp2Address = new Address(); emp2Address.AddressLine1 = "Gates Foundation"; emp2Address.AddressLine2 = "One Microsoft Way"; emp2Address.City = "Redmond"; emp2Address.State = "WA"; emp2Address.Zip = "98052-6399"; var emp2 = new Employee(); emp2.FirstName = "Melinda"; emp2.LastName = "Gates"; emp2.EmployeeAddress = emp2Address; So when I pass these two employee objects to my GetDifferingProperties method currently it returns FirstName and EmployeeAddress, but it does not tell me which exact property (which in this case is Address1) in the EmployeeAddress has changed. How can I tweak this method to get something like EmployeeAddress.Address1?

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