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Secure Your Wireless Router: 8 Things You Can Do Right Now

- by Chris Hoffman

A security researcher recently discovered a backdoor in many D-Link routers, allowing anyone to access the router without knowing the username or password. This isn’t the first router security issue and won’t be the last. To protect yourself, you should ensure that your router is configured securely. This is about more than just enabling Wi-Fi encryption and not hosting an open Wi-Fi network. Disable Remote Access Routers offer a web interface, allowing you to configure them through a browser. The router runs a web server and makes this web page available when you’re on the router’s local network. However, most routers offer a “remote access” feature that allows you to access this web interface from anywhere in the world. Even if you set a username and password, if you have a D-Link router affected by this vulnerability, anyone would be able to log in without any credentials. If you have remote access disabled, you’d be safe from people remotely accessing your router and tampering with it. To do this, open your router’s web interface and look for the “Remote Access,” “Remote Administration,” or “Remote Management” feature. Ensure it’s disabled — it should be disabled by default on most routers, but it’s good to check. Update the Firmware Like our operating systems, web browsers, and every other piece of software we use, router software isn’t perfect. The router’s firmware — essentially the software running on the router — may have security flaws. Router manufacturers may release firmware updates that fix such security holes, although they quickly discontinue support for most routers and move on to the next models. Unfortunately, most routers don’t have an auto-update feature like Windows and our web browsers do — you have to check your router manufacturer’s website for a firmware update and install it manually via the router’s web interface. Check to be sure your router has the latest available firmware installed. Change Default Login Credentials Many routers have default login credentials that are fairly obvious, such as the password “admin”. If someone gained access to your router’s web interface through some sort of vulnerability or just by logging onto your Wi-Fi network, it would be easy to log in and tamper with the router’s settings. To avoid this, change the router’s password to a non-default password that an attacker couldn’t easily guess. Some routers even allow you to change the username you use to log into your router. Lock Down Wi-Fi Access If someone gains access to your Wi-Fi network, they could attempt to tamper with your router — or just do other bad things like snoop on your local file shares or use your connection to downloaded copyrighted content and get you in trouble. Running an open Wi-Fi network can be dangerous. To prevent this, ensure your router’s Wi-Fi is secure. This is pretty simple: Set it to use WPA2 encryption and use a reasonably secure passphrase. Don’t use the weaker WEP encryption or set an obvious passphrase like “password”. Disable UPnP A variety of UPnP flaws have been found in consumer routers. Tens of millions of consumer routers respond to UPnP requests from the Internet, allowing attackers on the Internet to remotely configure your router. Flash applets in your browser could use UPnP to open ports, making your computer more vulnerable. UPnP is fairly insecure for a variety of reasons. To avoid UPnP-based problems, disable UPnP on your router via its web interface. If you use software that needs ports forwarded — such as a BitTorrent client, game server, or communications program — you’ll have to forward ports on your router without relying on UPnP. Log Out of the Router’s Web Interface When You’re Done Configuring It Cross site scripting (XSS) flaws have been found in some routers. A router with such an XSS flaw could be controlled by a malicious web page, allowing the web page to configure settings while you’re logged in. If your router is using its default username and password, it would be easy for the malicious web page to gain access. Even if you changed your router’s password, it would be theoretically possible for a website to use your logged-in session to access your router and modify its settings. To prevent this, just log out of your router when you’re done configuring it — if you can’t do that, you may want to clear your browser cookies. This isn’t something to be too paranoid about, but logging out of your router when you’re done using it is a quick and easy thing to do. Change the Router’s Local IP Address If you’re really paranoid, you may be able to change your router’s local IP address. For example, if its default address is 192.168.0.1, you could change it to 192.168.0.150. If the router itself were vulnerable and some sort of malicious script in your web browser attempted to exploit a cross site scripting vulnerability, accessing known-vulnerable routers at their local IP address and tampering with them, the attack would fail. This step isn’t completely necessary, especially since it wouldn’t protect against local attackers — if someone were on your network or software was running on your PC, they’d be able to determine your router’s IP address and connect to it. Install Third-Party Firmwares If you’re really worried about security, you could also install a third-party firmware such as DD-WRT or OpenWRT. You won’t find obscure back doors added by the router’s manufacturer in these alternative firmwares. Consumer routers are shaping up to be a perfect storm of security problems — they’re not automatically updated with new security patches, they’re connected directly to the Internet, manufacturers quickly stop supporting them, and many consumer routers seem to be full of bad code that leads to UPnP exploits and easy-to-exploit backdoors. It’s smart to take some basic precautions. Image Credit: Nuscreen on Flickr

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“It’s only test code…”

- by Chris George

“Let me hack this in, it’s only test code”, “Don’t worry about getting it reviewed, it’s only test code”, “It doesn’t have to be elegant or efficient, it’s only test code”… do these phrases sound familiar? Chances are if you’ve working with test automation, at one point or other you will have heard these phrases, you have probably even used them yourself! What is certain is that code written under this “it’s only test code” mantra will come back and bite you in the arse! I’ve recently encountered a case where a test was giving a false positive, therefore hiding a real product bug because that test code was very badly written. Firstly it was very difficult to understand what the test was actually trying to achieve let alone how it was doing it, and this complexity masked a simple logic error. These issues are real and they do happen. Let’s take a step back from this and look at what we are trying to do. We are writing test code that tests product code, and we do this to create a suite of tests that will help protect our software against regressions. This test code is making sure that the product behaves as it should by employing some sort of expected result verification. The simple cases of these are generally not a problem. However, automation allows us to explore more complex scenarios in many more permutations. As this complexity increases then so does the complexity of the test code. It is at this point that code which has not been architected properly will cause problems. Keep your friends close… So, how do we make sure we are doing it right? The development teams I have worked on have always had Test Engineers working very closely with their Software Engineers. This is something that I have always tried to take full advantage of. They are coding experts! So run your ideas past them, ask for advice on how to structure your code, help you design your data structures. This may require a shift in your teams viewpoint, as contrary to this section title and folklore, Software Engineers are not actually the mortal enemy of Test Engineers. As time progresses, and test automation becomes more and more ingrained in what we do, the two roles are converging more than ever. Over the 16 years I have spent as a Test Engineer, I have seen the grey area between the two roles grow significantly larger. This serves to strengthen the relationship and common bond between the two roles which helps to make test code activities so much easier! Pair for the win Possibly the best thing you could do to write good test code is to pair program on the task. This will serve a few purposes. you will get the benefit of the Software Engineers knowledge and experience the Software Engineer will gain knowledge on the testing process. Sharing the love is a wonderful thing! two pairs of eyes are always better than one… And so are two brains. Between the two of you, I will guarantee you will derive more useful test cases than if it was just one of you. Code reviews Another policy which certainly pays dividends is the practice of code reviews. By having one of your peers review your code before you commit it serves two purposes. Firstly, it forces you to explain your code. Just the act of doing this will often pick up errors in your code. Secondly, it gets yet another pair of eyes on your code! I cannot stress enough how important code reviews are. The benefits they offer apply as much to product code as test code. In short, Software and Test Engineers should all be doing them! It can be extended even further by getting test code reviewed by a Software Engineer and a Test Engineer, and likewise product code. This serves to keep both functions in the loop with changes going on within your code base. Learn from your devs I briefly touched on this earlier but I’d like to go into more detail here. Pairing with your Software Engineers when writing your test code is such an amazing opportunity to improve your coding skills. As I sit here writing this article waiting to be called into court for jury service, it reminds me that it takes a lot of patience to be a Test Engineer, almost as much as it takes to be a juror! However tempting it is to go rushing in and start writing your automated tests, resist that urge. Discuss what you want to achieve then talk through the approach you’re going to take. Then code it up together. I find it really enlightening to ask questions like ‘is there a better way to do this?’ Or ‘is this how you would code it?’ The latter question, especially, is where I learn the most. I’ve found that most Software Engineers will be reluctant to show you the ‘right way’ to code something when writing tests because they perceive the ‘right way’ to be too complicated for the Test Engineer (e.g. not mentioning LINQ and instead doing something verbose). So by asking how THEY would code it, it unleashes their true dev-ness and advanced code usually ensues! I would like to point out, however, that you don’t have to accept their method as the final answer. On numerous occasions I have opted for the more simple/verbose solution because I found the code written by the Software Engineer too advanced and therefore I would find it unreadable when I return to the code in a months’ time! Always keep the target audience in mind when writing clever code, and in my case that is mostly Test Engineers.

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ANTS CLR and Memory Profiler In Depth Review (Part 2 of 2 – Memory Profiler)

- by ToStringTheory

One of the things that people might not know about me, is my obsession to make my code as efficient as possible. Many people might not realize how much of a task or undertaking that this might be, but it is surely a task as monumental as climbing Mount Everest, except this time it is a challenge for the mind… In trying to make code efficient, there are many different factors that play a part – size of project or solution, tiers, language used, experience and training of the programmer, technologies used, maintainability of the code – the list can go on for quite some time. I spend quite a bit of time when developing trying to determine what is the best way to implement a feature to accomplish the efficiency that I look to achieve. One program that I have recently come to learn about – Red Gate ANTS Performance (CLR) and Memory profiler gives me tools to accomplish that job more efficiently as well. In this review, I am going to cover some of the features of the ANTS memory profiler set by compiling some hideous example code to test against. Notice As a member of the Geeks With Blogs Influencers program, one of the perks is the ability to review products, in exchange for a free license to the program. I have not let this affect my opinions of the product in any way, and Red Gate nor Geeks With Blogs has tried to influence my opinion regarding this product in any way. Introduction – Part 2 In my last post, I reviewed the feature packed Red Gate ANTS Performance Profiler. Separate from the Red Gate Performance Profiler is the Red Gate ANTS Memory Profiler – a simple, easy to use utility for checking how your application is handling memory management… A tool that I wish I had had many times in the past. This post will be focusing on the ANTS Memory Profiler and its tool set. The memory profiler has a large assortment of features just like the Performance Profiler, with the new session looking nearly exactly alike: ANTS Memory Profiler Memory profiling is not something that I have to do very often… In the past, the few cases I’ve had to find a memory leak in an application I have usually just had to trace the code of the operations being performed to look for oddities… Sadly, I have come across more undisposed/non-using’ed IDisposable objects, usually from ADO.Net than I would like to ever see. Support is not fun, however using ANTS Memory Profiler makes this task easier. For this round of testing, I am going to use the same code from my previous example, using the WPF application. This time, I will choose the ‘Profile Memory’ option from the ANTS menu in Visual Studio, which launches the solution in its currently configured state/start-up project, and then launches the ANTS Memory Profiler to help. It prepopulates all of the fields with the current project information, and all I have to do is select the ‘Start Profiling’ option. When the window comes up, it is actually quite barren, just giving ideas on how to work the profiler. You start by getting to the point in your application that you want to profile, and then taking a ‘Memory Snapshot’. This performs a full garbage collection, and snapshots the managed heap. Using the same WPF app as before, I will go ahead and take a snapshot now. As you can see, ANTS is already giving me lots of information regarding the snapshot, however this is just a snapshot. The whole point of the profiler is to perform an action, usually one where a memory problem is being noticed, and then take another snapshot and perform a diff between them to see what has changed. I am going to go ahead and generate 5000 primes, and then take another snapshot: As you can see, ANTS is already giving me a lot of new information about this snapshot compared to the last. Information such as difference in memory usage, fragmentation, class usage, etc… If you take more snapshots, you can use the dropdown at the top to set your actual comparison snapshots. If you beneath the timeline, you will see a breadcrumb trail showing how best to approach profiling memory using ANTS. When you first do the comparison, you start on the Summary screen. You can either use the charts at the bottom, or switch to the class list screen to get to the next step. Here is the class list screen: As you can see, it lists information about all of the instances between the snapshots, as well as at the bottom giving you a way to filter by telling ANTS what your problem is. I am going to go ahead and select the Int16[] to look at the Instance Categorizer Using the instance categorizer, you can travel backwards to see where all of the instances are coming from. It may be hard to see in this image, but hopefully the lightbox (click on it) will help: I can see that all of these instances are rooted to the application through the UI TextBlock control. This image will probably be even harder to see, however using the ‘Instance Retention Graph’, you can trace an objects memory inheritance up the chain to see its roots as well. This is a simple example, as this is simply a known element. Usually you would be profiling an actual problem, and comparing those differences. I know in the past, I have spotted a problem where a new context was created per page load, and it was rooted into the application through an event. As the application began to grow, performance and reliability problems started to emerge. A tool like this would have been a great way to identify the problem quickly. Overview Overall, I think that the Red Gate ANTS Memory Profiler is a great utility for debugging those pesky leaks. 3 Biggest Pros: Easy to use interface with lots of options for configuring profiling session Intuitive and helpful interface for drilling down from summary, to instance, to root graphs ANTS provides an API for controlling the profiler. Not many options, but still helpful. 2 Biggest Cons: Inability to automatically snapshot the memory by interval Lack of complete integration with Visual Studio via an extension panel Ratings Ease of Use (9/10) – I really do believe that they have brought simplicity to the once difficult task of memory profiling. I especially liked how it stepped you further into the drilldown by directing you towards the best options. Effectiveness (10/10) – I believe that the profiler does EXACTLY what it purports to do. Features (7/10) – A really great set of features all around in the application, however, I would like to see some ability for automatically triggering snapshots based on intervals or framework level items such as events. Customer Service (10/10) – My entire experience with Red Gate personnel has been nothing but good. their people are friendly, helpful, and happy! UI / UX (9/10) – The interface is very easy to get around, and all of the options are easy to find. With a little bit of poking around, you’ll be optimizing Hello World in no time flat! Overall (9/10) – Overall, I am happy with the Memory Profiler and its features, as well as with the service I received when working with the Red Gate personnel. Thank you for reading up to here, or skipping ahead – I told you it would be shorter! Please, if you do try the product, drop me a message and let me know what you think! I would love to hear any opinions you may have on the product. Code Feel free to download the code I used above – download via DropBox

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Introduction to Human Workflow 11g

- by agiovannetti

Human Workflow is a component of SOA Suite just like BPEL, Mediator, Business Rules, etc. The Human Workflow component allows you to incorporate human intervention in a business process. You can use Human Workflow to create a business process that requires a manager to approve purchase orders greater than $10,000; or a business process that handles article reviews in which a group of reviewers need to vote/approve an article before it gets published. Human Workflow can handle the task assignment and routing as well as the generation of notifications to the participants. There are three common patterns or usages of Human Workflow: 1) Approval Scenarios: manage documents and other transactional data through approval chains . For example: approve expense report, vacation approval, hiring approval, etc. 2) Reviews by multiple users or groups: group collaboration and review of documents or proposals. For example, processing a sales quote which is subject to review by multiple people. 3) Case Management: workflows around work management or case management. For example, processing a service request. This could be routed to various people who all need to modify the task. It may also incorporate ad hoc routing which is unknown at design time. SOA 11g Human Workflow includes the following features: Assignment and routing of tasks to the correct users or groups. Deadlines, escalations, notifications, and other features required for ensuring the timely performance of a task. Presentation of tasks to end users through a variety of mechanisms, including a Worklist application. Organization, filtering, prioritization and other features required for end users to productively perform their tasks. Reports, reassignments, load balancing and other features required by supervisors and business owners to manage the performance of tasks. Human Workflow Architecture The Human Workflow component is divided into 3 modules: the service interface, the task definition and the client interface module. The Service Interface handles the interaction with BPEL and other components. The Client Interface handles the presentation of task data through clients like the Worklist application, portals and notification channels. The task definition module is in charge of managing the lifecycle of a task. Who should get the task assigned? What should happen next with the task? When must the task be completed? Should the task be escalated?, etc Stages and Participants When you create a Human Task you need to specify how the task is assigned and routed. The first step is to define the stages and participants. A stage is just a logical group. A participant can be a user, a group of users or an application role. The participants indicate the type of assignment and routing that will be performed. Stages can be sequential or in parallel. You can combine them to create any usage you require. See diagram below: Assignment and Routing There are different ways a task can be assigned and routed: Single Approver: task is assigned to a single user, group or role. For example, a vacation request is assigned to a manager. If the manager approves or rejects the request, the employee is notified with the decision. If the task is assigned to a group then once one of managers acts on it, the task is completed. Parallel : task is assigned to a set of people that must work in parallel. This is commonly used for voting. For example, a task gets approved once 50% of the participants approve it. You can also set it up to be a unanimous vote. Serial : participants must work in sequence. The most common scenario for this is management chain escalation. FYI (For Your Information) : task is assigned to participants who can view it, add comments and attachments, but can not modify or complete the task. Task Actions The following is the list of actions that can be performed on a task: Claim : if a task is assigned to a group or multiple users, then the task must be claimed first to be able to act on it. Escalate : if the participant is not able to complete a task, he/she can escalate it. The task is reassigned to his/her manager (up one level in a hierarchy). Pushback : the task is sent back to the previous assignee. Reassign :if the participant is a manager, he/she can delegate a task to his/her reports. Release : if a task is assigned to a group or multiple users, it can be released if the user who claimed the task cannot complete the task. Any of the other assignees can claim and complete the task. Request Information and Submit Information : use when the participant needs to supply more information or to request more information from the task creator or any of the previous assignees. Suspend and Resume :if a task is not relevant, it can be suspended. A suspension is indefinite. It does not expire until Resume is used to resume working on the task. Withdraw : if the creator of a task does not want to continue with it, for example, he wants to cancel a vacation request, he can withdraw the task. The business process determines what happens next. Renew : if a task is about to expire, the participant can renew it. The task expiration date is extended one week. Notifications Human Workflow provides a mechanism for sending notifications to participants to alert them of changes on a task. Notifications can be sent via email, telephone voice message, instant messaging (IM) or short message service (SMS). Notifications can be sent when the task status changes to any of the following: Assigned/renewed/delegated/reassigned/escalated Completed Error Expired Request Info Resume Suspended Added/Updated comments and/or attachments Updated Outcome Withdraw Other Actions (e.g. acquiring a task) Here is an example of an email notification: Worklist Application Oracle BPM Worklist application is the default user interface included in SOA Suite. It allows users to access and act on tasks that have been assigned to them. For example, from the Worklist application, a loan agent can review loan applications or a manager can approve employee vacation requests. Through the Worklist Application users can: Perform authorized actions on tasks, acquire and check out shared tasks, define personal to-do tasks and define subtasks. Filter tasks view based on various criteria. Work with standard work queues, such as high priority tasks, tasks due soon and so on. Work queues allow users to create a custom view to group a subset of tasks in the worklist, for example, high priority tasks, tasks due in 24 hours, expense approval tasks and more. Define custom work queues. Gain proxy access to part of another user's tasks. Define custom vacation rules and delegation rules. Enable group owners to define task dispatching rules for shared tasks. Collect a complete workflow history and audit trail. Use digital signatures for tasks. Run reports like Unattended tasks, Tasks productivity, etc. Here is a screenshoot of what the Worklist Application looks like. On the right hand side you can see the tasks that have been assigned to the user and the task's detail. References Introduction to SOA Suite 11g Human Workflow Webcast Note 1452937.2 Human Workflow Information Center Using the Human Workflow Service Component 11.1.1.6 Human Workflow Samples Human Workflow APIs Java Docs

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Restoring databases to a set drive and directory

- by okeofs

Restoring databases to a set drive and directory Introduction Often people say that necessity is the mother of invention. In this case I was faced with the dilemma of having to restore several databases, with multiple ‘ndf’ files, and having to restore them with different physical file names, drives and directories on servers other than the servers from which they originated. As most of us would do, I went to Google to see if I could find some code to achieve this task and found some interesting snippets on Pinal Dave’s website. Naturally, I had to take it further than the code snippet, HOWEVER it was a great place to start. Creating a temp table to hold database file details First off, I created a temp table which would hold the details of the individual data files within the database. Although there are a plethora of fields (within the temp table below), I utilize LogicalName only within this example. The temporary table structure may be seen below: create table #tmp ( LogicalName nvarchar(128) ,PhysicalName nvarchar(260) ,Type char(1) ,FileGroupName nvarchar(128) ,Size numeric(20,0) ,MaxSize numeric(20,0), Fileid tinyint, CreateLSN numeric(25,0), DropLSN numeric(25, 0), UniqueID uniqueidentifier, ReadOnlyLSN numeric(25,0), ReadWriteLSN numeric(25,0), BackupSizeInBytes bigint, SourceBlocSize int, FileGroupId int, LogGroupGUID uniqueidentifier, DifferentialBaseLSN numeric(25,0), DifferentialBaseGUID uniqueidentifier, IsReadOnly bit, IsPresent bit, TDEThumbPrint varchar(50) ) We now declare and populate a variable(@path), setting the variable to the path to our SOURCE database backup. declare @path varchar(50) set @path = 'P:\DATA\MYDATABASE.bak' From this point, we insert the file details of our database into the temp table. Note that we do so by utilizing a restore statement HOWEVER doing so in ‘filelistonly’ mode. insert #tmp EXEC ('restore filelistonly from disk = ''' + @path + '''') At this point, I depart from what I gleaned from Pinal Dave. I now instantiate a few more local variables. The use of each variable will be evident within the cursor (which follows): Declare @RestoreString as Varchar(max) Declare @NRestoreString as NVarchar(max) Declare @LogicalName as varchar(75) Declare @counter as int Declare @rows as int set @counter = 1 select @rows = COUNT(*) from #tmp -- Count the number of records in the temp -- table Declaring and populating the cursor At this point I do realize that many people are cringing about the use of a cursor. Being an Oracle professional as well, I have learnt that there is a time and place for cursors. I would remind the reader that the data that will be read into the cursor is from a local temp table and as such, any locking of the records (within the temp table) is not really an issue. DECLARE MY_CURSOR Cursor FOR Select LogicalName From #tmp Parsing the logical names from within the cursor. A small caveat that works in our favour, is that the first logical name (of our database) is the logical name of the primary data file (.mdf). Other files, except for the very last logical name, belong to secondary data files. The last logical name is that of our database log file. I now open my cursor and populate the variable @RestoreString Open My_Cursor set @RestoreString = 'RESTORE DATABASE [MYDATABASE] FROM DISK = N''P:\DATA\ MYDATABASE.bak''' + ' with ' We now fetch the first record from the temp table. Fetch NEXT FROM MY_Cursor INTO @LogicalName While there are STILL records left within the cursor, we dynamically build our restore string. Note that we are using concatenation to create ‘one big restore executable string’. Note also that the target physical file name is hardwired, as is the target directory. While (@@FETCH_STATUS <> -1) BEGIN IF (@@FETCH_STATUS <> -2) -- As long as there are no rows missing select @RestoreString = case when @counter = 1 then -- This is the mdf file @RestoreString + 'move N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.mdf' + '''' + ', ' -- OK, if it passes through here we are dealing with an .ndf file -- Note that Counter must be greater than 1 and less than the number of rows. when @counter > 1 and @counter < @rows then -- These are the ndf file(s) @RestoreString + 'move N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.ndf' + '''' + ', ' -- OK, if it passes through here we are dealing with the log file When @LogicalName like '%log%' then @RestoreString + 'move N''' + @LogicalName + '''' + ' TO N’’X:\DATA1\'+ @LogicalName + '.ldf' +'''' end --Increment the counter set @counter = @counter + 1 FETCH NEXT FROM MY_CURSOR INTO @LogicalName END At this point we have populated the varchar(max) variable @RestoreString with a concatenation of all the necessary file names. What we now need to do is to run the sp_executesql stored procedure, to effect the restore. First, we must place our ‘concatenated string’ into an nvarchar based variable. Obviously this will only work as long as the length of @RestoreString is less than varchar(max) / 2. set @NRestoreString = @RestoreString EXEC sp_executesql @NRestoreString Upon completion of this step, the database should be restored to the server. I now close and deallocate the cursor, and to be clean, I would also drop my temp table. CLOSE MY_CURSOR DEALLOCATE MY_CURSOR GO Conclusion Restoration of databases on different servers with different physical names and on different drives are a fact of life. Through the use of a few variables and a simple cursor, we may achieve an efficient and effective way to achieve this task.

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How to begin? Windows 8 Development

- by Dennis Vroegop

Ok. I convinced you in my last post to do some Win8 development. You want a piece of that cake, or whatever your reasons may be. Good! Welcome to the club! Now let me ask you a question: what are you going to write? Ah. That’s the big one, isn’t it? What indeed? If you have been creating applications for computers before you’re in for quite a shock. The way people perceive apps on a tablet is quite different from what we know as applications. There’s a reason we call them apps instead of applications! Yes, technically they are applications but we don’t call them apps only because it sounds cool. The abbreviated form of the word applications itself is a pointer. Apps are small. Apps are focused. Apps are more lightweight. Apps do one thing but they do that one thing extremely good. In the ‘old’ days we wrote huge systems. We build ecosystems of services, screens, databases and more to create a system that provides value for the user. Think about it: what application do you use most at work? Can you in one sentence describe what it is, or what it does and yet still distinctively describe its purpose? I doubt you can. Let’s have a look at Outlouk. We all know it and we all love or hate it. But what is it? A mail program? No, there’s so much more there: calendar, contacts, RSS feeds and so on. Some call it a ‘collaboration’ application but that’s not really true as well. After all, why should a collaboration application give me my schedule for the day? I think the best way to describe Outlook is “client for Exchange” although that isn’t accurate either. Anyway: Outlook is a great application but it’s not an ‘app’ and therefor not very suitable for WinRT. Ok. Disclaimer here: yes, you can write big applications for WinRT. Some will. But that’s not what 99.9% of the developers will do. So I am stating here that big applications are not meant for WinRT. If 0.01% of the developers think that this is nonsense then they are welcome to go ahead but for the majority here this is not what we’re talking about. So: Apps are small, lightweight and good at what they do but only at that. If you’re a Phone developer you already know that: Phone apps on any platform fit the description I have above. If you’ve ever worked in a large cooperation before you might have seen one of these before: the Mission Statement. It’s supposed to be a oneliner that sums up what the company is supposed to do. Funny enough: although this doesn’t work for large companies it does work for defining your app. A mission statement for an app describes what it does. If it doesn’t fit in the mission statement then your app is going to get to big and will fail. A statement like this should be in the following style “<your app name> is the best app to <describe single task>” Fill in the blanks, write it and go! Mmm.. not really. There are some things there we need to think about. But the statement is a very, very important one. If you cannot fit your app in that line you’re preparing to fail. Your app will become to big, its purpose will be unclear and it will be hard to use. People won’t download it and those who do will give it a bad rating therefor preventing that huge success you’ve been dreaming about. Stick to the statement! Ok, let’s give it a try: “PlanesAreCool” is the best app to do planespotting in the field. You might have seen these people along runways of airports: taking photographs of airplanes and noting down their numbers and arrival- and departure times. We are going to help them out with our great app! If you look at the statement, can you guess what it does? I bet you can. If you find out it isn’t clear enough of if it’s too broad, refine it. This is probably the most important step in the development of your app so give it enough time! So. We’ve got the statement. Print it out, stick it to the wall and look at it. What does it tell you? If you see this, what do you think the app does? Write that down. Sit down with some friends and talk about it. What do they expect from an app like this? Write that down as well. Brainstorm. Make a list of features. This is mine: Note planes Look up aircraft carriers Add pictures of that plane Look up airfields Notify friends of new spots Look up details of a type of plane Plot a graph with arrival and departure times Share new spots on social media Look up history of a particular aircraft Compare your spots with friends Write down arrival times Write down departure times Write down wind conditions Write down the runway they take Look up weather conditions for next spotting day Invite friends to join you for a day of spotting. Now, I must make it clear that I am not a planespotter nor do I know what one does. So if the above list makes no sense, I apologize. There is a lesson: write apps for stuff you know about…. First of all, let’s look at our statement and then go through the list of features. Remove everything that has nothing to do with that statement! If you end up with an empty list, try again with both steps. Note planes Look up aircraft carriers Add pictures of that plane Look up airfields Notify friends of new spots Look up details of a type of plane Plot a graph with arrival and departure times Share new spots on social media Look up history of a particular aircraft Compare your spots with friends Write down arrival times Write down departure times Write down wind conditions Write down the runway they take Look up weather conditions for next spotting day Invite friends to join you for a day of spotting. That's better. The things I removed could be pretty useful to a plane spotter and could be fun to write. But do they match the statement? I said that the app is for spotting in the field, so “look up airfields” doesn’t belong there: I know where I am so why look it up? And the same goes for inviting friends or looking up the weather conditions for tomorrow. I am at the airfield right now, looking through my binoculars at the planes. I know the weather now and I don’t care about tomorrow. If you feel the items you’ve crossed out are valuable, then why not write another app? One that says “SpotNoter” is the best app for preparing a day of spotting with my friends. That’s a different app! Remember: Win8 apps are small and very good at doing ONE thing, and one thing only! If you have made that list, it’s time to prepare the navigation of your app. The navigation is how users see your app and how they use it. We’ll do that next time!

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Box2D blocky map. Body, Fixtures a huge map and performance

- by Solom

Right now I'm still in the planning phase of a my very first game. I'm creating a "Minecraft"-like game in 2D that features blocks that can be destroyed as well as players moving around the map. For creating the map I chose a 2D-Array of Integers that represent the Block ID. For testing purposes I created a huge map (16348 * 256) and in my prototype that didn't use Box2D everything worked like a charm. I only rendered those blocks that where within the bounds of my camera and got 60 fps straight. The problem started when I decided to use an existing physics-solution rather than implementing my own one. What I had was basically simple hitboxes around the blocks and then I had to manually check if the player collided with any of those in his neighborhood. For more advanced physics as well as the collision detection I want to switch over to Box2D. The problem I have right now is ... how to go about the bodies? I mean, the blocks are of a static bodytype. They don't move on their own, they just are there to be collided with. But as far as I can see it, every block needs his own body with a rectangular fixture attached to it, so as to be destroyable. But for a huge map such as mine, this turns out to be a real performance bottle-neck. (In fact even a rather small map [compared to the other] of 1024*256 is unplayable.) I mean I create thousands of thousands of blocks. Even if I just render those that are in my immediate neighborhood there are hundreds of them and (at least with the debugRenderer) I drop to 1 fps really quickly (on my own "monster machine"). I thought about strategies like creating just one body, attaching multiple fixtures and only if a fixture got hit, separate it from the body, create a new one and destroy it, but this didn't turn out quite as successful as hoped. (In fact the core just dumps. Ah hello C! I really missed you :X) Here is the code: public class Box2DGameScreen implements Screen { private World world; private Box2DDebugRenderer debugRenderer; private OrthographicCamera camera; private final float TIMESTEP = 1 / 60f; // 1/60 of a second -> 1 frame per second private final int VELOCITYITERATIONS = 8; private final int POSITIONITERATIONS = 3; private Map map; private BodyDef blockBodyDef; private FixtureDef blockFixtureDef; private BodyDef groundDef; private Body ground; private PolygonShape rectangleShape; @Override public void show() { world = new World(new Vector2(0, -9.81f), true); debugRenderer = new Box2DDebugRenderer(); camera = new OrthographicCamera(); // Pixel:Meter = 16:1 // Body definition BodyDef ballDef = new BodyDef(); ballDef.type = BodyDef.BodyType.DynamicBody; ballDef.position.set(0, 1); // Fixture definition FixtureDef ballFixtureDef = new FixtureDef(); ballFixtureDef.shape = new CircleShape(); ballFixtureDef.shape.setRadius(.5f); // 0,5 meter ballFixtureDef.restitution = 0.75f; // between 0 (not jumping up at all) and 1 (jumping up the same amount as it fell down) ballFixtureDef.density = 2.5f; // kg / m² ballFixtureDef.friction = 0.25f; // between 0 (sliding like ice) and 1 (not sliding) // world.createBody(ballDef).createFixture(ballFixtureDef); groundDef = new BodyDef(); groundDef.type = BodyDef.BodyType.StaticBody; groundDef.position.set(0, 0); ground = world.createBody(groundDef); this.map = new Map(20, 20); rectangleShape = new PolygonShape(); // rectangleShape.setAsBox(1, 1); blockFixtureDef = new FixtureDef(); // blockFixtureDef.shape = rectangleShape; blockFixtureDef.restitution = 0.1f; blockFixtureDef.density = 10f; blockFixtureDef.friction = 0.9f; } @Override public void render(float delta) { Gdx.gl.glClearColor(1, 1, 1, 1); Gdx.gl.glClear(GL20.GL_COLOR_BUFFER_BIT); debugRenderer.render(world, camera.combined); drawMap(); world.step(TIMESTEP, VELOCITYITERATIONS, POSITIONITERATIONS); } private void drawMap() { for(int a = 0; a < map.getHeight(); a++) { /* if(camera.position.y - (camera.viewportHeight/2) > a) continue; if(camera.position.y - (camera.viewportHeight/2) < a) break; */ for(int b = 0; b < map.getWidth(); b++) { /* if(camera.position.x - (camera.viewportWidth/2) > b) continue; if(camera.position.x - (camera.viewportWidth/2) < b) break; */ /* blockBodyDef = new BodyDef(); blockBodyDef.type = BodyDef.BodyType.StaticBody; blockBodyDef.position.set(b, a); world.createBody(blockBodyDef).createFixture(blockFixtureDef); */ PolygonShape rectangleShape = new PolygonShape(); rectangleShape.setAsBox(1, 1, new Vector2(b, a), 0); blockFixtureDef.shape = rectangleShape; ground.createFixture(blockFixtureDef); rectangleShape.dispose(); } } } @Override public void resize(int width, int height) { camera.viewportWidth = width / 16; camera.viewportHeight = height / 16; camera.update(); } @Override public void hide() { dispose(); } @Override public void pause() { } @Override public void resume() { } @Override public void dispose() { world.dispose(); debugRenderer.dispose(); } } As you can see I'm facing multiple problems here. I'm not quite sure how to check for the bounds but also if the map is bigger than 24*24 like 1024*256 Java just crashes -.-. And with 24*24 I get like 9 fps. So I'm doing something really terrible here, it seems and I assume that there most be a (much more performant) way, even with Box2D's awesome physics. Any other ideas? Thanks in advance!

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Combining Shared Secret and Certificates

- by Michael Stephenson

As discussed in the introduction article this walkthrough will explain how you can implement WCF security with the Windows Azure Service Bus to ensure that you can protect your endpoint in the cloud with a shared secret but also combine this with certificates so that you can identify the sender of the message. Prerequisites As in the previous article before going into the walk through I want to explain a few assumptions about the scenario we are implementing but to keep the article shorter I am not going to walk through all of the steps in how to setup some of this. In the solution we have a simple console application which will represent the client application. There is also the services WCF application which contains the WCF service we will expose via the Windows Azure Service Bus. The WCF Service application in this example was hosted in IIS 7 on Windows 2008 R2 with AppFabric Server installed and configured to auto-start the WCF listening services. I am not going to go through significant detail around the IIS setup because it should not matter in relation to this article however if you want to understand more about how to configure WCF and IIS for such a scenario please refer to the following paper which goes into a lot of detail about how to configure this. The link is: http://tinyurl.com/8s5nwrz Setting up the Certificates To keep the post and sample simple I am going to use the local computer store for all certificates but this bit is really just the same as setting up certificates for an example where you are using WCF without using Windows Azure Service Bus. In the sample I have included two batch files which you can use to create the sample certificates or remove them. Basically you will end up with: A certificate called PocServerCert in the personal store for the local computer which will be used by the WCF Service component A certificate called PocClientCert in the personal store for the local computer which will be used by the client application A root certificate in the Root store called PocRootCA with its associated revocation list which is the root from which the client and server certificates were created For the sample Im just using development certificates like you would normally, and you can see exactly how these are configured and placed in the stores from the batch files in the solution using makecert and certmgr. The Service Component To begin with let's look at the service component and how it can be configured to listen to the service bus using a shared secret but to also accept a username token from the client. In the sample the service component is called Acme.Azure.ServiceBus.Poc.Cert.Services. It has a single service which is the Visual Studio template for a WCF service when you add a new WCF Service Application so we have a service called Service1 with its Echo method. Nothing special so far!.... The next step is to look at the web.config file to see how we have configured the WCF service. In the services section of the WCF configuration you can see I have created my service and I have created a local endpoint which I simply used to do a little bit of diagnostics and to check it was working, but more importantly there is the Windows Azure endpoint which is using the ws2007HttpRelayBinding (note that this should also work just the same if your using netTcpRelayBinding). The key points to note on the above picture are the service behavior called MyServiceBehaviour and the service bus endpoints behavior called MyEndpointBehaviour. We will go into these in more detail later. The Relay Binding The relay binding for the service has been configured to use the TransportWithMessageCredential security mode. This is the important bit where the transport security really relates to the interaction between the service and listening to the Azure Service Bus and the message credential is where we will use our certificate like we have specified in the message/clientCrentialType attribute. Note also that we have left the relayClientAuthenticationType set to RelayAccessToken. This means that authentication will be made against ACS for accessing the service bus and messages will not be accepted from any sender who has not been authenticated by ACS. The Endpoint Behaviour In the below picture you can see the endpoint behavior which is configured to use the shared secret client credential for accessing the service bus and also for diagnostic purposes I have included the service registry element. Hopefully if you are familiar with using Windows Azure Service Bus relay feature the above is very familiar to you and this is a very common setup for this section. There is nothing specific to the username token implementation here. The Service Behaviour Now we come to the bit with most of the certificate stuff in it. When you configure the service behavior I have included the serviceCredentials element and then setup to use the clientCertificate check and also specifying the serviceCertificate with information on how to find the servers certificate in the store. I have also added a serviceAuthorization section where I will implement my own authorization component to perform additional security checks after the service has validated that the message was signed with a good certificate. I also have the same serviceSecurityAudit configuration to log access to my service. My Authorization Manager The below picture shows you implementation of my authorization manager. WCF will eventually hand off the message to my authorization component before it calls the service code. This is where I can perform some logic to check if the identity is allowed to access resources. In this case I am simple rejecting messages from anyone except the PocClientCertificate. The Client Now let's take a look at the client side of this solution and how we can configure the client to authenticate against ACS but also send a certificate over to the service component so it can implement additional security checks on-premise. I have a console application and in the program class I want to use the proxy generated with Add Service Reference to send a message via the Azure Service Bus. You can see in my WCF client configuration below I have setup my details for the azure service bus url and am using the ws2007HttpRelayBinding. Next is my configuration for the relay binding. You can see below I have configured security to use TransportWithMessageCredential so we will flow the token from a certificate with the message and also the RelayAccessToken relayClientAuthenticationType which means the component will validate against ACS before being allowed to access the relay endpoint to send a message. After the binding we need to configure the endpoint behavior like in the below picture. This contains the normal transportClientEndpointBehaviour to setup the ACS shared secret configuration but we have also configured the clientCertificate to look for the PocClientCert. Finally below we have the code of the client in the console application which will call the service bus. You can see that we have created our proxy and then made a normal call to a WCF in exactly the normal way but the configuration will jump in and ensure that a token is passed representing the client certificate. Conclusion As you can see from the above walkthrough it is not too difficult to configure a service to use both a shared secret and certificate based token at the same time. This gives you the power and protection offered by the access control service in the cloud but also the ability to flow additional tokens to the on-premise component for additional security features to be implemented. Sample The sample used in this post is available at the following location: https://s3.amazonaws.com/CSCBlogSamples/Acme.Azure.ServiceBus.Poc.Cert.zip

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Different Not Automatically Implies Better

- by Alois Kraus

Originally posted on: http://geekswithblogs.net/akraus1/archive/2013/11/05/154556.aspxRecently I was digging deeper why some WCF hosted workflow application did consume quite a lot of memory although it did basically only load a xaml workflow. The first tool of choice is Process Explorer or even better Process Hacker (has more options and the best feature copy&paste does work). The three most important numbers of a process with regards to memory are Working Set, Private Working Set and Private Bytes. Working set is the currently consumed physical memory (parts can be shared between processes e.g. loaded dlls which are read only) Private Working Set is the physical memory needed by this process which is not shareable Private Bytes is the number of non shareable which is only visible in the current process (e.g. all new, malloc, VirtualAlloc calls do create private bytes) When you have a bigger workflow it can consume under 64 bit easily 500MB for a 1-2 MB xaml file. This does not look very scalable. Under 64 bit the issue is excessive private bytes consumption and not the managed heap. The picture is quite different for 32 bit which looks a bit strange but it seems that the hosted VB compiler is a lot less memory hungry under 32 bit. I did try to repro the issue with a medium sized xaml file (400KB) which does contain 1000 variables and 1000 if which can be represented by C# code like this: string Var1; string Var2; ... string Var1000; if (!String.IsNullOrEmpty(Var1) ) { Console.WriteLine(“Var1”); } if (!String.IsNullOrEmpty(Var2) ) { Console.WriteLine(“Var2”); } .... Since WF is based on VB.NET expressions you are bound to the hosted VB.NET compiler which does result in (x64) 140 MB of private bytes which is ca. 140 KB for each if clause which is quite a lot if you think about the actually present functionality. But there is hope. .NET 4.5 does allow now C# expressions for WF which is a major step forward for all C# lovers. I did create some simple patcher to “cross compile” my xaml to C# expressions. Lets look at the result: C# Expressions VB Expressions x86 x86 On my home machine I have only 32 bit which gives you quite exactly half of the memory consumption under 64 bit. C# expressions are 10 times more memory hungry than VB.NET expressions! I wanted to do more with less memory but instead it did consume a magnitude more memory. That is surprising to say the least. The workflow does initialize in about the same time under x64 and x86 where the VB code does it in 2s whereas the C# version needs 18s. Also nearly ten times slower. That is a too high price to pay for any bigger sized xaml workflow to convert from VB.NET to C# expressions. If I do reduce the number of expressions to 500 then it does need 400MB which is about half of the memory. It seems that the cost per if does rise linear with the number of total expressions in a xaml workflow. Expression Language Cost per IF Startup Time C# 1000 Ifs x64 1,5 MB 18s C# 500 Ifs x64 750 KB 9s VB 1000 Ifs x64 140 KB 2s VB 500 Ifs x64 70 KB 1s Now we can directly compare two MS implementations. It is clear that the VB.NET compiler uses the same underlying structure but it has much higher offset compared to the highly inefficient C# expression compiler. I have filed a connect bug here with a harsher wording about recent advances in memory consumption. The funniest thing is that one MS employee did give an Azure AppFabric demo around early 2011 which was so slow that he needed to investigate with xperf. He was after startup time and the call stacks with regards to VB.NET expression compilation were remarkably similar. In fact I only found this post by googling for parts of my call stacks. … “C# expressions will be coming soon to WF, and that will have different performance characteristics than VB” … What did he know Jan 2011 what I did no know until today? ;-). He knew that C# expression will come but that they will not be automatically have better footprint. It is about time to fix that. In its current state C# expressions are not usable for bigger workflows. That also explains the headline for today. You can cheat startup time by prestarting workflows so that the demo looks nice and snappy but it does hurt scalability a lot since you do need much more memory than necessary. I did find the stacks by enabling virtual allocation tracking within XPerf which is still the best tool out there. But first you need to look at your process to check where the memory is hiding: For the C# Expression compiler you do not need xperf. You can directly dump the managed heap and check with a profiler of your choice. But if the allocations are happening on the Private Data ( VirtualAlloc ) you can find it with xperf. There is a nice video on channel 9 explaining VirtualAlloc tracking it in greater detail. If your data allocations are on the Heap it does mean that the C/C++ runtime did create a heap for you where all malloc, new calls do allocate from it. You can enable heap tracing with xperf and full call stack support as well which is doable via xperf like it is shown also on channel 9. Or you can use WPRUI directly: To make “Heap Usage” it work you need to set for your executable the tracing flags (before you start it). For example devenv.exe HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\Image File Execution Options\devenv.exe DWORD TracingFlags 1 Do not forget to disable it after you did complete profiling the process or it will impact the startup time quite a lot. You can with xperf attach directly to a running process and collect heap allocation information from a gone wild process. Very handy if you need to find out what a process was doing which has arrived in a funny state. “VirtualAlloc usage” does work without explicitly enabling stuff for a specific process and is always on machine wide. I had issues on my Windows 7 machines with the call stack collection and the latest Windows 8.1 Performance Toolkit. I was told that WPA from Windows 8.0 should work fine but I do not want to downgrade.

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Portable class libraries and fetching JSON

- by Jeff

After much delay, we finally have the Windows Phone 8 SDK to go along with the Windows 8 Store SDK, or whatever ridiculous name they’re giving it these days. (Seriously… that no one could come up with a suitable replacement for “metro” is disappointing in an otherwise exciting set of product launches.) One of the neat-o things is the potential for code reuse, particularly across Windows 8 and Windows Phone 8 apps. This is accomplished in part with portable class libraries, which allow you to share code between different types of projects. With some other techniques and quasi-hacks, you can share some amount of code, and I saw it mentioned in one of the Build videos that they’re seeing as much as 70% code reuse. Not bad. However, I’ve already hit a super annoying snag. It appears that the HttpClient class, with its idiot-proof async goodness, is not included in the Windows Phone 8 class libraries. Shock, gasp, horror, disappointment, etc. The delay in releasing it already caused dismay among developers, and I’m sure this won’t help. So I started refactoring some code I already had for a Windows 8 Store app (ugh) to accommodate the use of HttpWebRequest instead. I haven’t tried it in a Windows Phone 8 project beyond compiling, but it appears to work. I used this StackOverflow answer as a starting point since it’s been a long time since I used HttpWebRequest, and keep in mind that it has no exception handling. It needs refinement. The goal here is to new up the client, and call a method that returns some deserialized JSON objects from the Intertubes. Adding facilities for headers or cookies is probably a good next step. You need to use NuGet for a Json.NET reference. So here’s the start: using System.Net; using System.Threading.Tasks; using Newtonsoft.Json; using System.IO; namespace MahProject { public class ServiceClient<T> where T : class { public ServiceClient(string url) { _url = url; } private readonly string _url; public async Task<T> GetResult() { var response = await MakeAsyncRequest(_url); var result = JsonConvert.DeserializeObject<T>(response); return result; } public static Task<string> MakeAsyncRequest(string url) { var request = (HttpWebRequest)WebRequest.Create(url); request.ContentType = "application/json"; Task<WebResponse> task = Task.Factory.FromAsync( request.BeginGetResponse, asyncResult => request.EndGetResponse(asyncResult), null); return task.ContinueWith(t => ReadStreamFromResponse(t.Result)); } private static string ReadStreamFromResponse(WebResponse response) { using (var responseStream = response.GetResponseStream()) using (var reader = new StreamReader(responseStream)) { var content = reader.ReadToEnd(); return content; } } } } Calling it in some kind of repository class may look like this, if you wanted to return an array of Park objects (Park model class omitted because it doesn’t matter): public class ParkRepo { public async Task<Park[]> GetAllParks() { var client = new ServiceClient<Park[]>(http://superfoo/endpoint); return await client.GetResult(); } } And then from inside your WP8 or W8S app (see what I did there?), when you load state or do some kind of UI event handler (making sure the method uses the async keyword): var parkRepo = new ParkRepo(); var results = await parkRepo.GetAllParks(); // bind results to some UI or observable collection or something Hopefully this saves you a little time.

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Query optimization using composite indexes

- by xmarch

Many times, during the process of creating a new Coherence application, developers do not pay attention to the way cache queries are constructed; they only check that these queries comply with functional specs. Later, performance testing shows that these perform poorly and it is then when developers start working on improvements until the non-functional performance requirements are met. This post describes the optimization process of a real-life scenario, where using a composite attribute index has brought a radical improvement in query execution times. The execution times went down from 4 seconds to 2 milliseconds! E-commerce solution based on Oracle ATG – Endeca In the context of a new e-commerce solution based on Oracle ATG – Endeca, Oracle Coherence has been used to calculate and store SKU prices. In this architecture, a Coherence cache stores the final SKU prices used for Endeca baseline indexing. Each SKU price is calculated from a base SKU price and a series of calculations based on information from corporate global discounts. Corporate global discounts information is stored in an auxiliary Coherence cache with over 800.000 entries. In particular, to obtain each price the process needs to execute six queries over the global discount cache. After the implementation was finished, we discovered that the most expensive steps in the price calculation discount process were the global discounts cache query. This query has 10 parameters and is executed 6 times for each SKU price calculation. The steps taken to optimise this query are described below; Starting point Initial query was: String filter = "levelId = :iLevelId AND salesCompanyId = :iSalesCompanyId AND salesChannelId = :iSalesChannelId "+ "AND departmentId = :iDepartmentId AND familyId = :iFamilyId AND brand = :iBrand AND manufacturer = :iManufacturer "+ "AND areaId = :iAreaId AND endDate >= :iEndDate AND startDate <= :iStartDate"; Map<String, Object> params = new HashMap<String, Object>(10); // Fill all parameters. params.put("iLevelId", xxxx); // Executing filter. Filter globalDiscountsFilter = QueryHelper.createFilter(filter, params); NamedCache globalDiscountsCache = CacheFactory.getCache(CacheConstants.GLOBAL_DISCOUNTS_CACHE_NAME); Set applicableDiscounts = globalDiscountsCache.entrySet(globalDiscountsFilter); With the small dataset used for development the cache queries performed very well. However, when carrying out performance testing with a real-world sample size of 800,000 entries, each query execution was taking more than 4 seconds. First round of optimizations The first optimisation step was the creation of separate Coherence index for each of the 10 attributes used by the filter. This avoided object deserialization while executing the query. Each index was created as follows: globalDiscountsCache.addIndex(new ReflectionExtractor("getXXX" ) , false, null); After adding these indexes the query execution time was reduced to between 450 ms and 1s. However, these execution times were still not good enough. Second round of optimizations In this optimisation phase a Coherence query explain plan was used to identify how many entires each index reduced the results set by, along with the cost in ms of executing that part of the query. Though the explain plan showed that all the indexes for the query were being used, it also showed that the ordering of the query parameters was "sub-optimal". Parameters associated to object attributes with high-cardinality should appear at the beginning of the filter, or more specifically, the attributes that filters out the highest of number records should be placed at the beginning. But examining corporate global discount data we realized that depending on the values of the parameters used in the query the “good” order for the attributes was different. In particular, if the attributes brand and family had specific values it was more optimal to have a different query changing the order of the attributes. Ultimately, we ended up with three different optimal variants of the query that were used in its relevant cases: String filter = "brand = :iBrand AND familyId = :iFamilyId AND departmentId = :iDepartmentId AND levelId = :iLevelId "+ "AND manufacturer = :iManufacturer AND endDate >= :iEndDate AND salesCompanyId = :iSalesCompanyId "+ "AND areaId = :iAreaId AND salesChannelId = :iSalesChannelId AND startDate <= :iStartDate"; String filter = "familyId = :iFamilyId AND departmentId = :iDepartmentId AND levelId = :iLevelId AND brand = :iBrand "+ "AND manufacturer = :iManufacturer AND endDate >= :iEndDate AND salesCompanyId = :iSalesCompanyId "+ "AND areaId = :iAreaId AND salesChannelId = :iSalesChannelId AND startDate <= :iStartDate"; String filter = "brand = :iBrand AND departmentId = :iDepartmentId AND familyId = :iFamilyId AND levelId = :iLevelId "+ "AND manufacturer = :iManufacturer AND endDate >= :iEndDate AND salesCompanyId = :iSalesCompanyId "+ "AND areaId = :iAreaId AND salesChannelId = :iSalesChannelId AND startDate <= :iStartDate"; Using the appropriate query depending on the value of brand and family parameters the query execution time dropped to between 100 ms and 150 ms. But these these execution times were still not good enough and the solution was cumbersome. Third and last round of optimizations The third and final optimization was to introduce a composite index. However, this did mean that it was not possible to use the Coherence Query Language (CohQL), as composite indexes are not currently supporte in CohQL. As the original query had 8 parameters using EqualsFilter, 1 using GreaterEqualsFilter and 1 using LessEqualsFilter, the composite index was built for the 8 attributes using EqualsFilter. The final query had an EqualsFilter for the multiple extractor, a GreaterEqualsFilter and a LessEqualsFilter for the 2 remaining attributes. All individual indexes were dropped except the ones being used for LessEqualsFilter and GreaterEqualsFilter. We were now running in an scenario with an 8-attributes composite filter and 2 single attribute filters. The composite index created was as follows: ValueExtractor[] ve = { new ReflectionExtractor("getSalesChannelId" ), new ReflectionExtractor("getLevelId" ), new ReflectionExtractor("getAreaId" ), new ReflectionExtractor("getDepartmentId" ), new ReflectionExtractor("getFamilyId" ), new ReflectionExtractor("getManufacturer" ), new ReflectionExtractor("getBrand" ), new ReflectionExtractor("getSalesCompanyId" )}; MultiExtractor me = new MultiExtractor(ve); NamedCache globalDiscountsCache = CacheFactory.getCache(CacheConstants.GLOBAL_DISCOUNTS_CACHE_NAME); globalDiscountsCache.addIndex(me, false, null); And the final query was: ValueExtractor[] ve = { new ReflectionExtractor("getSalesChannelId" ), new ReflectionExtractor("getLevelId" ), new ReflectionExtractor("getAreaId" ), new ReflectionExtractor("getDepartmentId" ), new ReflectionExtractor("getFamilyId" ), new ReflectionExtractor("getManufacturer" ), new ReflectionExtractor("getBrand" ), new ReflectionExtractor("getSalesCompanyId" )}; MultiExtractor me = new MultiExtractor(ve); // Fill composite parameters.String SalesCompanyId = xxxx;...AndFilter composite = new AndFilter(new EqualsFilter(me, Arrays.asList(iSalesChannelId, iLevelId, iAreaId, iDepartmentId, iFamilyId, iManufacturer, iBrand, SalesCompanyId)), new GreaterEqualsFilter(new ReflectionExtractor("getEndDate" ), iEndDate)); AndFilter finalFilter = new AndFilter(composite, new LessEqualsFilter(new ReflectionExtractor("getStartDate" ), iStartDate)); NamedCache globalDiscountsCache = CacheFactory.getCache(CacheConstants.GLOBAL_DISCOUNTS_CACHE_NAME); Set applicableDiscounts = globalDiscountsCache.entrySet(finalFilter); Using this composite index the query improved dramatically and the execution time dropped to between 2 ms and 4 ms. These execution times completely met the non-functional performance requirements . It should be noticed than when using the composite index the order of the attributes inside the ValueExtractor was not relevant.

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Who could ask for more with LESS CSS? (Part 2 of 3–Setup)

- by ToStringTheory

Welcome to part two in my series covering the LESS CSS language. In the first post, I covered the two major CSS precompiled languages - LESS and SASS to a small extent, iterating over some of the features that you could expect to find in them. In this post, I will go a little further in depth into the setup and execution of using the LESS framework. Introduction It really doesn’t take too much to get LESS working in your project. The basic workflow will be including the necessary translator in your project, defining bundles for the LESS files, add the necessary code to your layouts.cshtml file, and finally add in all your necessary styles to the LESS files! Lets get started… New Project Just like all great experiments in Visual Studio, start up a File > New Project, and create a new MVC 4 Web Application. The Base Package After you have the new project spun up, use the Nuget Package Manager to install the Bundle Transformer: LESS package. This will take care of installing the main translator that we will be using for LESS code (dotless which is another Nuget package), as well as the core framework for the Bundle Transformer library. The installation will come up with some instructions in a readme file on how to modify your web.config to handle all your *.less requests through the Bundle Transformer, which passes the translating onto dotless. Where To Put These LESS Files?! This step isn’t really a requirement, however I find that I don’t like how ASP.Net MVC just has a content directory where they store CSS, content images, css images…. In my project, I went ahead and created a new directory just for styles – LESS files, CSS files, and images that are only referenced in LESS or CSS. Ignore the MVC directory as this was my testbed for another project I was working on at the same time. As you can see here, I have: A top level directory for images which contains only images used in a page A top level directory for scripts A top level directory for Styles A few directories for plugins I am using (Colrizr, JQueryUI, Farbtastic) Multiple *.less files for different functions (I’ll go over these in a minute) I find that this layout offers the best separation of content types. Bring Out Your Bundles! The next thing that we need to do is add in the necessary code for the bundling of these LESS files. Go ahead and open your BundleConfig.cs file, usually located in the /App_Start/ folder of the project. As you will see in a minute, instead of using the method Microsoft does in the base MVC 4 project, I change things up a bit. Define Constants The first thing I do is define constants for each of the virtual paths that will be used in the bundler: The main reason is that I hate magic strings in my program, so the fact that you first defined a virtual path in the BundleConfig file, and then used that path in the _Layout.cshtml file really irked me. Add Bundles to the BundleCollection Next, I am going to define the bundles for my styles in my AddStyleBundles method: That is all it takes to get all of my styles in play with LESS. The CssTransformer and NullOrderer types come from the Bundle Transformer we grabbed earlier. If we didn’t use that package, we would have to write our own function (not too hard, but why do it if it’s been done). I use the site.less file as my main hub for LESS - I will cover that more in the next section. Add Bundles To Layout.cshtml File With the constants in the BundleConfig file, instead of having to use the same magic string I defined for the bundle virtual path, I am able to do this: Notice here that besides the RenderSection magic strings (something I am working on in another side project), all of the bundles are now based on const strings. If I need to change the virtual path, I only have to do it in one place. Nifty! Get Started! We are now ready to roll! As I said in the previous section, I use the site.less file as a central hub for my styles: As seen here, I have a reset.css file which is a simple CSS reset. Next, I have created a file for managing all my color variables – colors.less: Here, you can see some of the standards I started to use, in this case for color variables. I define all color variables with the @col prefix. Currently, I am going for verbose variable names. The next file imported is my font.less file that defines the typeface information for the site: Simple enough. A couple of imports for fonts from Google, and then declaring variables for use throughout LESS. I also set up the heading sizes, margins, etc.. You can also see my current standardization for font declaration strings – @font. Next, I pull in a mixins.less file that I grabbed from the Twitter Bootstrap library that gives some useful parameterized mixins for use such as border-radius, gradient, box-shadow, etc… The common.less file is a file that just contains items that I will be defining that can be used across all my LESS files. Kind of like my own mixins or font-helpers: Finally I have my layout.less file that contains all of my definitions for general site layout – width, main/sidebar widths, footer layout, etc: That’s it! For the rest of my one off definitions/corrections, I am currently putting them into the site.less file beneath my original imports Note Probably my favorite side effect of using the LESS handler/translator while bundling is that it also does a CSS checkup when rendering… See, when your web.config is set to debug, bundling will output the url to the direct less file, not the bundle, and the http handler intercepts the call, compiles the less, and returns the result. If there is an error in your LESS code, the CSS file can be returned empty, or may have the error output as a comment on the first couple lines. If you have the web.config set to not debug, then if there is an error in your code, you will end up with the usual ASP.Net exception page (unless you catch the exception of course), with information regarding the failure of the conversion, such as brace mismatch, undefined variable, etc… I find it nifty. Conclusion This is really just the beginning. LESS is very powerful and exciting! My next post will show an actual working example of why LESS is so powerful with its functions and variables… At least I hope it will! As for now, if you have any questions, comments, or suggestions on my current practice, I would love to hear them! Feel free to drop a comment or shoot me an email using the contact page. In the mean time, I plan on posting the final post in this series tomorrow or the day after, with my side project, as well as a whole base ASP.Net MVC4 templated project with LESS added in it so that you can check out the layout I have in this post. Until next time…

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Java EE 7 Survey Results!

- by reza_rahman

On November 8th, the Java EE EG posted a survey to gather broad community feedback on a number of critical open issues. For reference, you can find the original survey here. We kept the survey open for about three weeks until November 30th. To our delight, over 1100 developers took time out of their busy lives to let their voices be heard! The results of the survey were sent to the EG on December 12th. The subsequent EG discussion is available here. The exact summary sent to the EG is available here. We would like to take this opportunity to thank each and every one the individuals who took the survey. It is very appreciated, encouraging and worth it's weight in gold. In particular, I tried to capture just some of the high-quality, intelligent, thoughtful and professional comments in the summary to the EG. I highly encourage you to continue to stay involved, perhaps through the Adopt-a-JSR program. We would also like to sincerely thank java.net, JavaLobby, TSS and InfoQ for helping spread the word about the survey. Below is a brief summary of the results... APIs to Add to Java EE 7 Full/Web Profile The first question asked which of the four new candidate APIs (WebSocket, JSON-P, JBatch and JCache) should be added to the Java EE 7 Full and Web profile respectively. As the following graph shows, there was significant support for adding all the new APIs to the full profile: Support is relatively the weakest for Batch 1.0, but still good. A lot of folks saw WebSocket 1.0 as a critical technology with comments such as this one: "A modern web application needs Web Sockets as first class citizens" While it is clearly seen as being important, a number of commenters expressed dissatisfaction with the lack of a higher-level JSON data binding API as illustrated by this comment: "How come we don't have a Data Binding API for JSON" JCache was also seen as being very important as expressed with comments like: "JCache should really be that foundational technology on which other specs have no fear to depend on" The results for the Web Profile is not surprising. While there is strong support for adding WebSocket 1.0 and JSON-P 1.0 to the Web Profile, support for adding JCache 1.0 and Batch 1.0 is relatively weak. There was actually significant opposition to adding Batch 1. 0 (with 51.8% casting a 'No' vote). Enabling CDI by Default The second question asked was whether CDI should be enabled in Java EE environments by default. A significant majority of 73.3% developers supported enabling CDI, only 13.8% opposed. Comments such as these two reflect a strong general support for CDI as well as a desire for better Java EE alignment with CDI: "CDI makes Java EE quite valuable!" "Would prefer to unify EJB, CDI and JSF lifecycles" There is, however, a palpable concern around the performance impact of enabling CDI by default as exemplified by this comment: "Java EE projects in most cases use CDI, hence it is sensible to enable CDI by default when creating a Java EE application. However, there are several issues if CDI is enabled by default: scanning can be slow - not all libs use CDI (hence, scanning is not needed)" Another significant concern appears to be around backwards compatibility and conflict with other JSR 330 implementations like Spring: "I am leaning towards yes, however can easily imagine situations where errors would be caused by automatically activating CDI, especially in cases of backward compatibility where another DI engine (such as Spring and the like) happens to use the same mechanics to inject dependencies and in that case there would be an overlap in injections and probably an uncertain outcome" Some commenters such as this one attempt to suggest solutions to these potential issues: "If you have Spring in use and use javax.inject.Inject then you might get some unexpected behavior that could be equally confusing. I guess there will be a way to switch CDI off. I'm tempted to say yes but am cautious for this reason" Consistent Usage of @Inject The third question was around using CDI/JSR 330 @Inject consistently vs. allowing JSRs to create their own injection annotations. A slight majority of 53.3% developers supported using @Inject consistently across JSRs. 28.8% said using custom injection annotations is OK, while 18.0% were not sure. The vast majority of commenters were strongly supportive of CDI and general Java EE alignment with CDI as illistrated by these comments: "Dependency Injection should be standard from now on in EE. It should use CDI as that is the DI mechanism in EE and is quite powerful. Having a new JSR specific DI mechanism to deal with just means more reflection, more proxies. JSRs should also be constructed to allow some of their objects Injectable. @Inject @TransactionalCache or @Inject @JMXBean etc...they should define the annotations and stereotypes to make their code less procedural. Dog food it. If there is a shortcoming in CDI for a JSR fix it and we will all be grateful" "We're trying to make this a comprehensive platform, right? Injection should be a fundamental part of the platform; everything else should build on the same common infrastructure. Each-having-their-own is just a recipe for chaos and having to learn the same thing 10 different ways" Expanding the Use of @Stereotype The fourth question was about expanding CDI @Stereotype to cover annotations across Java EE beyond just CDI. A significant majority of 62.3% developers supported expanding the use of @Stereotype, only 13.3% opposed. A majority of commenters supported the idea as well as the theme of general CDI/Java EE alignment as expressed in these examples: "Just like defining new types for (compositions of) existing classes, stereotypes can help make software development easier" "This is especially important if many EJB services are decoupled from the EJB component model and can be applied via individual annotations to Java EE components. @Stateless is a nicely compact annotation. Code will not improve if that will have to be applied in the future as @Transactional, @Pooled, @Secured, @Singlethreaded, @...." Some, however, expressed concerns around increased complexity such as this commenter: "Could be very convenient, but I'm afraid if it wouldn't make some important class annotations less visible" Expanding Interceptor Use The final set of questions was about expanding interceptors further across Java EE... A very solid 96.3% of developers wanted to expand interceptor use to all Java EE components. 35.7% even wanted to expand interceptors to other Java EE managed classes. Most developers (54.9%) were not sure if there is any place that injection is supported that should not support interceptors. 32.8% thought any place that supports injection should also support interceptors. Only 12.2% were certain that there are places where injection should be supported but not interceptors. The comments reflected the diversity of opinions, generally supportive of interceptors: "I think interceptors are as fundamental as injection and should be available anywhere in the platform" "The whole usage of interceptors still needs to take hold in Java programming, but it is a powerful technology that needs some time in the Sun. Basically it should become part of Java SE, maybe the next step after lambas?" A distinct chain of thought separated interceptors from filters and listeners: "I think that the Servlet API already provides a rich set of possibilities to hook yourself into different Servlet container events. I don't find a need to 'pollute' the Servlet model with the Interceptors API"

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Problem Implementing Texture on Libgdx Mesh of Randomized Terrain

- by BrotherJack

I'm having problems understanding how to apply a texture to a non-rectangular object. The following code creates textures such as this: from the debug renderer I think I've got the physical shape of the "earth" correct. However, I don't know how to apply a texture to it. I have a 50x50 pixel image (in the environment constructor as "dirt.png"), that I want to apply to the hills. I have a vague idea that this seems to involve the mesh class and possibly a ShapeRenderer, but the little i'm finding online is just confusing me. Bellow is code from the class that makes and regulates the terrain and the code in a separate file that is supposed to render it (but crashes on the mesh.render() call). Any pointers would be appreciated. public class Environment extends Actor{ Pixmap sky; public Texture groundTexture; Texture skyTexture; double tankypos; //TODO delete, temp public Tank etank; //TODO delete, temp int destructionRes; // how wide is a static pixel private final float viewWidth; private final float viewHeight; private ChainShape terrain; public Texture dirtTexture; private World world; public Mesh terrainMesh; private static final String LOG = Environment.class.getSimpleName(); // Constructor public Environment(Tank tank, FileHandle sfileHandle, float w, float h, int destructionRes) { world = new World(new Vector2(0, -10), true); this.destructionRes = destructionRes; sky = new Pixmap(sfileHandle); viewWidth = w; viewHeight = h; skyTexture = new Texture(sky); terrain = new ChainShape(); genTerrain((int)w, (int)h, 6); Texture tankSprite = new Texture(Gdx.files.internal("TankSpriteBase.png")); Texture turretSprite = new Texture(Gdx.files.internal("TankSpriteTurret.png")); tank = new Tank(0, true, tankSprite, turretSprite); Rectangle tankrect = new Rectangle(300, (int)tankypos, 44, 45); tank.setRect(tankrect); BodyDef terrainDef = new BodyDef(); terrainDef.type = BodyType.StaticBody; terrainDef.position.set(0, 0); Body terrainBody = world.createBody(terrainDef); FixtureDef fixtureDef = new FixtureDef(); fixtureDef.shape = terrain; terrainBody.createFixture(fixtureDef); BodyDef tankDef = new BodyDef(); Rectangle rect = tank.getRect(); tankDef.type = BodyType.DynamicBody; tankDef.position.set(0,0); tankDef.position.x = rect.x; tankDef.position.y = rect.y; Body tankBody = world.createBody(tankDef); FixtureDef tankFixture = new FixtureDef(); PolygonShape shape = new PolygonShape(); shape.setAsBox(rect.width*WORLD_TO_BOX, rect.height*WORLD_TO_BOX); fixtureDef.shape = shape; dirtTexture = new Texture(Gdx.files.internal("dirt.png")); etank = tank; } private void genTerrain(int w, int h, int hillnessFactor){ int width = w; int height = h; Random rand = new Random(); //min and max bracket the freq's of the sin/cos series //The higher the max the hillier the environment int min = 1; //allocating horizon for screen width Vector2[] horizon = new Vector2[width+2]; horizon[0] = new Vector2(0,0); double[] skyline = new double[width]; //TODO skyline necessary as an array? //ratio of amplitude of screen height to landscape variation double r = (int) 2.0/5.0; //number of terms to be used in sine/cosine series int n = 4; int[] f = new int[n*2]; //calculating omegas for sine series for(int i = 0; i < n*2 ; i ++){ f[i] = rand.nextInt(hillnessFactor - min + 1) + min; } //amp is the amplitude of the series int amp = (int) (r*height); double lastPoint = 0.0; for(int i = 0 ; i < width; i ++){ skyline[i] = 0; for(int j = 0; j < n; j++){ skyline[i] += ( Math.sin( (f[j]*Math.PI*i/height) ) + Math.cos(f[j+n]*Math.PI*i/height) ); } skyline[i] *= amp/(n*2); skyline[i] += (height/2); skyline[i] = (int)skyline[i]; //TODO Possible un-necessary float to int to float conversions tankypos = skyline[i]; horizon[i+1] = new Vector2((float)i, (float)skyline[i]); if(i == width) lastPoint = skyline[i]; } horizon[width+1] = new Vector2(800, (float)lastPoint); terrain.createChain(horizon); terrain.createLoop(horizon); //I have no idea if the following does anything useful :( terrainMesh = new Mesh(true, (width+2)*2, (width+2)*2, new VertexAttribute(Usage.Position, (width+2)*2, "a_position")); float[] vertices = new float[(width+2)*2]; short[] indices = new short[(width+2)*2]; for(int i=0; i < (width+2); i+=2){ vertices[i] = horizon[i].x; vertices[i+1] = horizon[i].y; indices[i] = (short)i; indices[i+1] = (short)(i+1); } terrainMesh.setVertices(vertices); terrainMesh.setIndices(indices); } Here is the code that is (supposed to) render the terrain. @Override public void render(float delta) { Gdx.gl.glClearColor(1, 1, 1, 1); Gdx.gl.glClear(GL10.GL_COLOR_BUFFER_BIT); // tell the camera to update its matrices. camera.update(); // tell the SpriteBatch to render in the // coordinate system specified by the camera. backgroundStage.draw(); backgroundStage.act(delta); uistage.draw(); uistage.act(delta); batch.begin(); debugRenderer.render(this.ground.getWorld(), camera.combined); batch.end(); //Gdx.graphics.getGL10().glEnable(GL10.GL_TEXTURE_2D); ground.dirtTexture.bind(); ground.terrainMesh.render(GL10.GL_TRIANGLE_FAN); //I'm particularly lost on this ground.step(); }

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Maintenance plans love story

- by Maria Zakourdaev

There are about 200 QA and DEV SQL Servers out there. There is a maintenance plan on many of them that performs a backup of all databases and removes the backup history files. First of all, I must admit that I’m no big fan of maintenance plans in particular or the SSIS packages in general. In this specific case, if I ever need to change anything in the way backup is performed, such as the compression feature or perform some other change, I have to open each plan one by one. This is quite a pain. Therefore, I have decided to replace the maintenance plans with a stored procedure that will perform exactly the same thing. Having such a procedure will allow me to open multiple server connections and just execute an ALTER PROCEDURE whenever I need to change anything in it. There is nothing like good ole T-SQL. The first challenge was to remove the unneeded maintenance plans. Of course, I didn’t want to do it server by server. I found the procedure msdb.dbo.sp_maintplan_delete_plan, but it only has a parameter for the maintenance plan id and it has no other parameters, like plan name, which would have been much more useful. Now I needed to find the table that holds all maintenance plans on the server. You would think that it would be msdb.dbo.sysdbmaintplans but, unfortunately, regardless of the number of maintenance plans on the instance, it contains just one row. After a while I found another table: msdb.dbo.sysmaintplan_subplans. It contains the plan id that I was looking for, in the plan_id column and well as the agent’s job id which is executing the plan’s package: That was all I needed and the rest turned out to be quite easy. Here is a script that can be executed against hundreds of servers from a multi-server query window to drop the specific maintenance plans. DECLARE @PlanID uniqueidentifier SELECT @PlanID = plan_id FROM msdb.dbo.sysmaintplan_subplans Where name like ‘BackupPlan%’ EXECUTE msdb.dbo.sp_maintplan_delete_plan @plan_id=@PlanID The second step was to create a procedure that will perform all of the old maintenance plan tasks: create a folder for each database, backup all databases on the server and clean up the old files. The script is below. Enjoy. ALTER PROCEDURE BackupAllDatabases @PrintMode BIT = 1 AS BEGIN DECLARE @BackupLocation VARCHAR(500) DECLARE @PurgeAferDays INT DECLARE @PurgingDate VARCHAR(30) DECLARE @SQLCmd VARCHAR(MAX) DECLARE @FileName VARCHAR(100) SET @PurgeAferDays = -14 SET @BackupLocation = '\\central_storage_servername\BACKUPS\'+@@servername SET @PurgingDate = CONVERT(VARCHAR(19), DATEADD (dd,@PurgeAferDays,GETDATE()),126) SET @FileName = '?_full_'+ + REPLACE(CONVERT(VARCHAR(19), GETDATE(),126),':','-') +'.bak'; SET @SQLCmd = ' IF ''?'' <> ''tempdb'' BEGIN EXECUTE master.dbo.xp_create_subdir N'''+@BackupLocation+'\?\'' ; BACKUP DATABASE ? TO DISK = N'''+@BackupLocation+'\?\'+@FileName+''' WITH NOFORMAT, NOINIT, SKIP, REWIND, NOUNLOAD, COMPRESSION, STATS = 10 ; EXECUTE master.dbo.xp_delete_file 0,N'''+@BackupLocation+'\?\'',N''bak'',N'''+@PurgingDate+''',1; END' IF @PrintMode = 1 BEGIN PRINT @SQLCmd END EXEC sp_MSforeachdb @SQLCmd END

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Increase moving speed of body

- by Siddharth

How to move ball speedily on the screen using box2d in libGDX? public class Box2DDemo implements ApplicationListener { private SpriteBatch batch; private TextureRegion texture; private World world; private Body groundDownBody, groundUpBody, groundLeftBody, groundRightBody, ballBody; private BodyDef groundBodyDef1, groundBodyDef2, groundBodyDef3, groundBodyDef4, ballBodyDef; private PolygonShape groundDownPoly, groundUpPoly, groundLeftPoly, groundRightPoly; private CircleShape ballPoly; private Sprite sprite; private FixtureDef fixtureDef; private Vector2 ballPosition; private Box2DDebugRenderer renderer; Vector2 vector2; @Override public void create() { texture = new TextureRegion(new Texture( Gdx.files.internal("img/red_ring.png"))); sprite = new Sprite(texture); sprite.setOrigin(sprite.getWidth() / 2, sprite.getHeight() / 2); batch = new SpriteBatch(); world = new World(new Vector2(0.0f, -10.0f), false); groundBodyDef1 = new BodyDef(); groundBodyDef1.type = BodyType.StaticBody; groundBodyDef1.position.x = 0.0f; groundBodyDef1.position.y = 0.0f; groundDownBody = world.createBody(groundBodyDef1); groundBodyDef2 = new BodyDef(); groundBodyDef2.type = BodyType.StaticBody; groundBodyDef2.position.x = 0f; groundBodyDef2.position.y = Gdx.graphics.getHeight(); groundUpBody = world.createBody(groundBodyDef2); groundBodyDef3 = new BodyDef(); groundBodyDef3.type = BodyType.StaticBody; groundBodyDef3.position.x = 0f; groundBodyDef3.position.y = 0f; groundLeftBody = world.createBody(groundBodyDef3); groundBodyDef4 = new BodyDef(); groundBodyDef4.type = BodyType.StaticBody; groundBodyDef4.position.x = Gdx.graphics.getWidth(); groundBodyDef4.position.y = 0f; groundRightBody = world.createBody(groundBodyDef4); groundDownPoly = new PolygonShape(); groundDownPoly.setAsBox(480.0f, 10f); fixtureDef = new FixtureDef(); fixtureDef.density = 0f; fixtureDef.restitution = 1f; fixtureDef.friction = 0f; fixtureDef.shape = groundDownPoly; fixtureDef.filter.groupIndex = 0; groundDownBody.createFixture(fixtureDef); groundUpPoly = new PolygonShape(); groundUpPoly.setAsBox(480.0f, 10f); fixtureDef = new FixtureDef(); fixtureDef.friction = 0f; fixtureDef.restitution = 0f; fixtureDef.density = 0f; fixtureDef.shape = groundUpPoly; fixtureDef.filter.groupIndex = 0; groundUpBody.createFixture(fixtureDef); groundLeftPoly = new PolygonShape(); groundLeftPoly.setAsBox(10f, 320f); fixtureDef = new FixtureDef(); fixtureDef.friction = 0f; fixtureDef.restitution = 0f; fixtureDef.density = 0f; fixtureDef.shape = groundLeftPoly; fixtureDef.filter.groupIndex = 0; groundLeftBody.createFixture(fixtureDef); groundRightPoly = new PolygonShape(); groundRightPoly.setAsBox(10f, 320f); fixtureDef = new FixtureDef(); fixtureDef.friction = 0f; fixtureDef.restitution = 0f; fixtureDef.density = 0f; fixtureDef.shape = groundRightPoly; fixtureDef.filter.groupIndex = 0; groundRightBody.createFixture(fixtureDef); ballPoly = new CircleShape(); ballPoly.setRadius(16f); fixtureDef = new FixtureDef(); fixtureDef.shape = ballPoly; fixtureDef.density = 1f; fixtureDef.friction = 1f; fixtureDef.restitution = 1f; ballBodyDef = new BodyDef(); ballBodyDef.type = BodyType.DynamicBody; ballBodyDef.position.x = (int) 200; ballBodyDef.position.y = (int) 200; ballBody = world.createBody(ballBodyDef); // ballBody.setLinearVelocity(200f, 200f); // ballBody.applyLinearImpulse(new Vector2(250f, 250f), // ballBody.getLocalCenter()); ballBody.createFixture(fixtureDef); renderer = new Box2DDebugRenderer(true, false, false); } @Override public void dispose() { ballPoly.dispose(); groundLeftPoly.dispose(); groundUpPoly.dispose(); groundDownPoly.dispose(); groundRightPoly.dispose(); world.destroyBody(ballBody); world.dispose(); } @Override public void pause() { } @Override public void render() { world.step(1f/30f, 3, 3); Gdx.gl.glClearColor(1f, 1f, 1f, 1f); Gdx.gl.glClear(GL10.GL_COLOR_BUFFER_BIT); batch.begin(); vector2 = ballBody.getLinearVelocity(); System.out.println("X=" + vector2.x + " Y=" + vector2.y); ballPosition = ballBody.getPosition(); renderer.render(world,batch.getProjectionMatrix()); // int preX = (int) (vector2.x / Math.abs(vector2.x)); // int preY = (int) (vector2.y / Math.abs(vector2.y)); // // if (Math.abs(vector2.x) == 0.0f) // ballBody1.setLinearVelocity(1.4142137f, vector2.y); // else if (Math.abs(vector2.x) < 1.4142137f) // ballBody1.setLinearVelocity(preX * 5, vector2.y); // // if (Math.abs(vector2.y) == 0.0f) // ballBody1.setLinearVelocity(vector2.x, 1.4142137f); // else if (Math.abs(vector2.y) < 1.4142137f) // ballBody1.setLinearVelocity(vector2.x, preY * 5); batch.draw(sprite, (ballPosition.x - (texture.getRegionWidth() / 2)), (ballPosition.y - (texture.getRegionHeight() / 2))); batch.end(); } @Override public void resize(int arg0, int arg1) { } @Override public void resume() { } } I implement above code but I can not achieve higher moving speed of the ball

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Rendering Flickr Cats Via Backbone.js

- by Geertjan

Create a JavaScript file and refer to it inside an HTML file. Then put this into the JavaScript file: (function($) { var CatCollection = Backbone.Collection.extend({ url: 'http://api.flickr.com/services/feeds/photos_public.gne?tags=cat&tagmode=any&format=json&jsoncallback=?', parse: function(response) { return response.items; } }); var CatView = Backbone.View.extend({ el: $('body'), initialize: function() { _.bindAll(this, 'render'); carCollectionInstance.fetch({ success: function(response, xhr) { catView.render(); } }); }, render: function() { $(this.el).append("<ul></ul>"); for (var i = 0; i < carCollectionInstance.length; i++) { $('ul', this.el).append("<li>" + i + carCollectionInstance.models[i].get("description") + "</li>"); } } }); var carCollectionInstance = new CatCollection(); var catView = new CatView(); })(jQuery); Apologies for any errors or misused idioms. It's my second day with Backbone.js, in fact, my second day with JavaScript. I haven't seen anywhere online so far where an example such as the above is found, though plenty that do kind of or pieces of the above, or explain in text, without an actual full example. The next step, and the only reason for the above experiment, is to create some JPA entities and expose them via RESTful webservices created on EJB methods, for consumption into an HTML5 application via a Backbone.js script very similar to the above.

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howto only tunnel specific hosts route through openvpn client on tomato

- by kcome

I am relatively newbie in networking world although I did coding and know some sysadmin background for a long time. and here I'm only one step from my destination. The whole picture is : at home I use one LinkSys E3000 as the gateway(don't know yet if this is it's name), wireless AP and no other routing/switching devices. It serves 1 PC and 1 Mac with LAN, 1 Mac Mini + 1 iPad + 2 smartphones with WIFI. My goal is use an openvpn client on the E3000 (with tomato firmware) and make my iPad and smartphone's all WiFi traffic through it, and other devices route remain the same non-openvpn route. So far I'm able to connect openvpn client on E3000 to an openvpn server, tunnel all my devices' all traffic through that openvpn connection. What's left is howto selectively route by source IP (at least in my guessing) to the tunnel while don't bother others. I had learned some 'iptables' and 'route' in past few days however without much luck, so here comes my question. Here are some info which will help you get the structure. ifconfig -a output, some useless lines striped, and in the web interface C0:C1:C0:1A:E0:28 is WAN, C0:C1:C0:1A:E0:27 is LAN, C0:C1:C0:1A:E0:29 is 2.4G wifi AP, C0:C1:C0:1A:E0:2A is 5G wifi AP. root@router:/tmp/home/root# ifconfig -a br0 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:27 inet addr:192.168.1.1 Bcast:192.168.1.255 Mask:255.255.255.0 UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 eth0 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:27 UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 eth1 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:29 UP BROADCAST RUNNING ALLMULTI MULTICAST MTU:1500 Metric:1 eth2 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:2A UP BROADCAST RUNNING ALLMULTI MULTICAST MTU:1500 Metric:1 lo Link encap:Local Loopback inet addr:127.0.0.1 Mask:255.0.0.0 inet6 addr: ::1/128 Scope:Host ppp0 Link encap:Point-to-Point Protocol inet addr:172.200.1.43 P-t-P:172.200.0.1 Mask:255.255.255.255 UP POINTOPOINT RUNNING MULTICAST MTU:1480 Metric:1 vlan1 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:27 UP BROADCAST RUNNING ALLMULTI MULTICAST MTU:1500 Metric:1 vlan2 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:28 UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 wl0.1 Link encap:Ethernet HWaddr C0:C1:C0:1A:E0:29 BROADCAST MULTICAST MTU:1500 Metric:1 brctl show output root@router:/tmp/home/root# brctl show bridge name bridge id STP enabled interfaces br0 8000.c0c1c01ae027 no vlan1 eth1 eth2 before openvpn route-up script root@router:/tmp/home/root# route -n Kernel IP routing table Destination Gateway Genmask Flags Metric Ref Use Iface 172.200.0.1 0.0.0.0 255.255.255.255 UH 0 0 0 ppp0 192.168.1.0 0.0.0.0 255.255.255.0 U 0 0 0 br0 127.0.0.0 0.0.0.0 255.0.0.0 U 0 0 0 lo 0.0.0.0 172.200.0.1 0.0.0.0 UG 0 0 0 ppp0 openvpn server push PUSH: Received control message: 'PUSH_REPLY,redirect-gateway,dhcp-option DNS 8.8.8.8,route 172.20.0.1,topology net30,ping 10,ping-restart 120,ifconfig 172.20.0.6 172.20.0.5' openvpn's stock route-up script Apr 24 14:52:06 router daemon.notice openvpn[1768]: /sbin/ifconfig tun11 172.20.0.6 pointopoint 172.20.0.5 mtu 1500 Apr 24 14:52:08 router daemon.notice openvpn[1768]: /sbin/route add -net 72.14.177.29 netmask 255.255.255.255 gw 172.200.0.1 Apr 24 14:52:08 router daemon.notice openvpn[1768]: /sbin/route add -net 0.0.0.0 netmask 128.0.0.0 gw 172.20.0.5 Apr 24 14:52:08 router daemon.notice openvpn[1768]: /sbin/route add -net 128.0.0.0 netmask 128.0.0.0 gw 172.20.0.5 Apr 24 14:52:08 router daemon.notice openvpn[1768]: /sbin/route add -net 172.20.0.1 netmask 255.255.255.255 gw 172.20.0.5 route after openvpn root@router:/tmp/home/root# route -n Kernel IP routing table Destination Gateway Genmask Flags Metric Ref Use Iface 172.20.0.5 0.0.0.0 255.255.255.255 UH 0 0 0 tun11 72.14.177.29 172.200.0.1 255.255.255.255 UGH 0 0 0 ppp0 172.200.0.1 0.0.0.0 255.255.255.255 UH 0 0 0 ppp0 172.20.0.1 172.20.0.5 255.255.255.255 UGH 0 0 0 tun11 192.168.1.0 0.0.0.0 255.255.255.0 U 0 0 0 br0 127.0.0.0 0.0.0.0 255.0.0.0 U 0 0 0 lo 0.0.0.0 172.20.0.5 128.0.0.0 UG 0 0 0 tun11 128.0.0.0 172.20.0.5 128.0.0.0 UG 0 0 0 tun11 0.0.0.0 172.200.0.1 0.0.0.0 UG 0 0 0 ppp0 something I had noticed and tried: * on the web interface of openvpn client there is an option "Create NAT on tunnel", if i check this, there is the following script (probably executed after openvpn connection established) root@router:/tmp/home/root# cat /tmp/etc/openvpn/fw/client1-fw.sh #!/bin/sh iptables -I INPUT -i tun11 -j ACCEPT iptables -I FORWARD -i tun11 -j ACCEPT iptables -t nat -I POSTROUTING -s 192.168.1.0/255.255.255.0 -o tun11 -j MASQUERADE if i uncheck this option, the last line will not appear. Then I guess probably the my issue will be solved by iptables and NAT related commands, I just haven't got enough knowledge to figure them out. I tried run iptables -t nat -I POSTROUTING -s 192.168.1.6 -o tun11 -j MASQUERADE manually after openvpn connected (192.168.1.6 is the ip address of my iPad), then my iPad get internet with openvpn tunnel, however all other devices can't reach internet. in case if needed, here is the iptables about NAT root@router:/tmp/home/root# iptables -t nat -L -n Chain PREROUTING (policy ACCEPT) target prot opt source destination DROP all -- 0.0.0.0/0 192.168.1.0/24 WANPREROUTING all -- 0.0.0.0/0 172.200.1.43 upnp all -- 0.0.0.0/0 172.200.1.43 Chain POSTROUTING (policy ACCEPT) target prot opt source destination MASQUERADE all -- 0.0.0.0/0 0.0.0.0/0 SNAT all -- 192.168.1.0/24 192.168.1.0/24 to:192.168.1.1 Chain OUTPUT (policy ACCEPT) target prot opt source destination Chain WANPREROUTING (1 references) target prot opt source destination DNAT icmp -- 0.0.0.0/0 0.0.0.0/0 to:192.168.1.1 Chain upnp (1 references) target prot opt source destination DNAT udp -- 0.0.0.0/0 0.0.0.0/0 udp dpt:5353 to:192.168.1.3:5353 Thanks in advance for helping and read this so much, I hope i made every info you need to give a help :)

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Postfix configuration - Uing virtual min but server is bouncing back my mail.

- by brodiebrodie

I have no experience in setting up postfix, and thought virtualmin minght do the legwork for me. Appears not. When I try to send mail to the domain (either [email protected] [email protected] or [email protected]) I get the following message returned This is the mail system at host dedq239.localdomain. I'm sorry to have to inform you that your message could not be delivered to one or more recipients. It's attached below. For further assistance, please send mail to <postmaster> If you do so, please include this problem report. You can delete your own text from the attached returned message. The mail system <[email protected]> (expanded from <[email protected]>): User unknown in virtual alias table Final-Recipient: rfc822; [email protected] Original-Recipient: rfc822;[email protected] Action: failed Status: 5.0.0 Diagnostic-Code: X-Postfix; User unknown in virtual alias table How can I diagnose the problem here? It seems that the mail gets to my server but the server fails to locally deliver the message to the correct user. (This is a guess, truthfully I have no idea what is happening). I have checked my virtual alias table and it seems to be set up correctly (I can post if this would be helpful). Can anyone give me a clue as to the next step? Thanks alias_database = hash:/etc/aliases alias_maps = hash:/etc/aliases broken_sasl_auth_clients = yes command_directory = /usr/sbin config_directory = /etc/postfix daemon_directory = /usr/libexec/postfix debug_peer_level = 2 html_directory = no local_recipient_maps = $virtual_mailbox_maps mailq_path = /usr/bin/mailq.postfix manpage_directory = /usr/share/man mydestination = $myhostname, localhost.$mydomain, localhost, $mydomain myorigin = $mydomain newaliases_path = /usr/bin/newaliases.postfix readme_directory = /usr/share/doc/postfix-2.3.3/README_FILES sample_directory = /usr/share/doc/postfix-2.3.3/samples sendmail_path = /usr/sbin/sendmail.postfix setgid_group = postdrop smtpd_recipient_restrictions = permit_mynetworks reject_unauth_destination smtpd_sasl_auth_enable = yes soft_bounce = no unknown_local_recipient_reject_code = 550 virtual_alias_maps = hash:/etc/postfix/virtual My mail log file (the last entry) Sep 30 15:13:47 dedq239 postfix/cleanup[7237]: 207C6B18158: message-id=<[email protected]> Sep 30 15:13:47 dedq239 postfix/qmgr[7177]: 207C6B18158: from=<[email protected]>, size=1805, nrcpt=1 (queue active) Sep 30 15:13:47 dedq239 postfix/error[7238]: 207C6B18158: to=<[email protected]>, orig_to=<[email protected]>, relay=none, delay=0.64, delays=0.61/0.01/0/0.02, dsn=5.0.0, status=bounced (User unknown in virtual alias table) Sep 30 15:13:47 dedq239 postfix/cleanup[7237]: 8DC13B18169: message-id=<[email protected]> Sep 30 15:13:47 dedq239 postfix/qmgr[7177]: 8DC13B18169: from=<>, size=3691, nrcpt=1 (queue active) Sep 30 15:13:47 dedq239 postfix/bounce[7239]: 207C6B18158: sender non-delivery notification: 8DC13B18169 Sep 30 15:13:47 dedq239 postfix/qmgr[7177]: 207C6B18158: removed Sep 30 15:13:48 dedq239 postfix/smtp[7240]: 8DC13B18169: to=<[email protected]>, relay=gmail-smtp-in.l.google.com[209.85.216.55]:25, delay=1.3, delays=0.02/0.01/0.58/0.75, dsn=2.0.0, status=sent (250 2.0.0 OK 1254348828 36si15082901pxi.91) Sep 30 15:13:48 dedq239 postfix/qmgr[7177]: 8DC13B18169: removed Sep 30 15:14:17 dedq239 postfix/smtpd[7233]: disconnect from mail-bw0-f228.google.com[209.85.218.228] etc.aliases file below I have not touched this file - myvirtualdomain is a replacement for my real domain name # Aliases in this file will NOT be expanded in the header from # Mail, but WILL be visible over networks or from /bin/mail. # # >>>>>>>>>> The program "newaliases" must be run after # >> NOTE >> this file is updated for any changes to # >>>>>>>>>> show through to sendmail. # # Basic system aliases -- these MUST be present. mailer-daemon: postmaster postmaster: root # General redirections for pseudo accounts. bin: root daemon: root adm: root lp: root sync: root shutdown: root halt: root mail: root news: root uucp: root operator: root games: root gopher: root ftp: root nobody: root radiusd: root nut: root dbus: root vcsa: root canna: root wnn: root rpm: root nscd: root pcap: root apache: root webalizer: root dovecot: root fax: root quagga: root radvd: root pvm: root amanda: root privoxy: root ident: root named: root xfs: root gdm: root mailnull: root postgres: root sshd: root smmsp: root postfix: root netdump: root ldap: root squid: root ntp: root mysql: root desktop: root rpcuser: root rpc: root nfsnobody: root ingres: root system: root toor: root manager: root dumper: root abuse: root newsadm: news newsadmin: news usenet: news ftpadm: ftp ftpadmin: ftp ftp-adm: ftp ftp-admin: ftp www: webmaster webmaster: root noc: root security: root hostmaster: root info: postmaster marketing: postmaster sales: postmaster support: postmaster # trap decode to catch security attacks decode: root # Person who should get root's mail #root: marc abuse-myvirtualdomain.com: [email protected] My etc/postfix/virtual file is below - again myvirtualdomain is a replacement. I think this file was generated by Virtualmin and I have tried messing around with is with no success... This is the version without my changes. myunixusername@myvirtualdomain .com myunixusername myvirtualdomain .com myvirtualdomain.com [email protected] [email protected] [email protected] [email protected] [email protected] [email protected] [email protected] [email protected]

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Arch Linux with an nginx/django setup refuses to display ANYTHING

- by Holland

I'm on Amazon Ec2, with an Arch Linux server. While I truly am loving it, I'm having the issue of actually getting nginx to display anything. Everytime I try to throw my hostname into the browser, the browser states that it's not available for some reason - almost as if the host doesn't even exist. One thing I'd like to know is, how can I get this up and running? Is there a specific arch linux configuration I have to do to make it web accessible? I have port 80 open, as well as port 22. I've tried using gunicorn, python-flup, and nginx. Nginx Config user http; worker_processes 1; #error_log logs/error.log; #error_log logs/error.log notice; #error_log logs/error.log info; pid logs/nginx.pid; events { worker_connections 1024; } http { include mime.types; default_type application/octet-stream; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; access_log logs/access.log main; sendfile on; #tcp_nopush on; #keepalive_timeout 0; keepalive_timeout 65; #gzip on; server { listen 80; server_name _; access_log /var/log/nginx/access.log; error_log /var/log/nginx/error.log; #charset koi8-r; location ^~ /media/ { root /path/to/media; } location ^~ /admin-media/ { root /usr/lib/python2.7/site-packages/django/contrib/admin/media; } location / { root /path/to/root/; fastcgi_pass 127.0.0.1:8080; fastcgi_param SERVER_NAME $server_name; fastcgi_param SERVER_PORT $server_port; fastcgi_param SERVER_PROTOCOL $server_protocol; fastcgi_param PATH_INFO $fastcgi_script_name; fastcgi_param REQUEST_METHOD $request_method; fastcgi_param QUERY_STRING $query_string; fastcgi_param CONTENT_TYPE $content_type; fastcgi_param CONTENT_LENGTH $content_length; fastcgi_pass_header Authorization; fastcgi_intercept_errors off; fastcgi_index index.html; index index.htm index.html; } error_page 500 502 503 504 /50x.html; location = /50x.html { root /etc/nginx/html/50x.html; } } # server { # listen 80; # server_name localhost; #charset koi8-r; #access_log logs/host.access.log main; # location / { # root html; # index index.html index.htm; # } #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # #error_page 500 502 503 504 /50x.html; #location = /50x.html { root html; #} # proxy the PHP scripts to Apache listening on 127.0.0.1:80 # #location ~ \.php$ { # proxy_pass http://127.0.0.1; #} # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000 # #location ~ \.php$ { # root html; # fastcgi_pass 127.0.0.1:9000; # fastcgi_index index.php; # fastcgi_param SCRIPT_FILENAME /scripts$fastcgi_script_name; # include fastcgi_params; #} # deny access to .htaccess files, if Apache's document root # concurs with nginx's one # #location ~ /\.ht { # deny all; #} #} # another virtual host using mix of IP-, name-, and port-based configuration # #server { # listen 8000; # listen somename:8080; # server_name somename alias another.alias; # location / { # root html; # index index.html index.htm; # } #} # HTTPS server # #server { # listen 443; # server_name localhost; # ssl on; # ssl_certificate cert.pem; # ssl_certificate_key cert.key; # ssl_session_timeout 5m; # ssl_protocols SSLv2 SSLv3 TLSv1; # ssl_ciphers HIGH:!aNULL:!MD5; # ssl_prefer_server_ciphers on; # location / { # root html; # index index.html index.htm; # } #} } I can't quite tell if it's a server issue or a configuration issue: I've followed so many guides now I can't even count them all. The thing is that Django itself is working fine, and my permissions to the document root of the where the site files are stored is 777. Ontop of that, I have a git repo which works perfectly fine, and django, python, and runfcgi all start without issues. The same goes for gunicorn, when I do a gunicorn_django -b 0.0.0.0:8000 in my document root. Here is my output from that: 2012-04-15 05:17:37 [3124] [INFO] Starting gunicorn 0.14.2 2012-04-15 05:17:37 [3124] [INFO] Listening at: http://0.0.0.0:8081 (3124) 2012-04-15 05:17:37 [3124] [INFO] Using worker: sync 2012-04-15 05:17:37 [3127] [INFO] Booting worker with pid: 3127 As far as I know, everything seems fine, as well as error.log and access.log for nginx. The access log is completely blank, for that matter. I just feel lost here; what would be a step in the right direction to bebugging an issue such as this?

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Windows installation repair option not showing up

- by Carl

I'm trying to repair an existing Windows XP installation. Following the instructions from http://www.microsoft.com/windowsxp/using/helpandsupport/learnmore/tips/doug92.mspx this should work: When the Press any key to boot from CD message is displayed on your screen, press a key to start your computer from the Windows XP CD. Press ENTER when you see the message To setup Windows XP now, and then press ENTER displayed on the Welcome to Setup screen. Do not choose the option to press R to use the Recovery Console. In the Windows XP Licensing Agreement, press F8 to agree to the license agreement. Make sure that your current installation of Windows XP is selected in the box, and then press R to repair Windows XP. Follow the instructions on the screen to complete Setup. On step 5 pressing R does nothing and there is nothing on the screen saying it would. When I just select to install I get a message that a previous installation is there and proceeding will destroy it and installed applications, I can optionally select a directory other than c:\windows, and I can optionally format before continuing. I had tried to go from SP2-SP3. It failed, and then I couldn't get to Safe Mode. I put the SP1 disk back in to do a repair, and I don't see that option. (I don't have an SP2 boot/install disk, I just have the non-boot upgrade package.) UPDATE: Upon loading the Recovery Console, I get a message saying The system registry does not appear to have an active ControlSet key. The system registry may be damaged. You can try restarting it with the Last Known Good configuration or you can try repairing the installation of Windows using the setup program's repair and recovery options. I then did bootcfg /scan - "successful" ... Total installs: 1 ... [1] c:\windows - with the c:\windows command prompt below it. bootcfg /list gives [1] Windows XP Pro; OS Load Options /noexecute=optin /fastdetect; OS Location: c:\windows I followed the instructions at http://michaelstevenstech.com/XPrepairinstall.htm - "Warning 2" link copy E:\i386\ntldr C:\ copy E:\i386\ntdetect.com C:\ attrib -h -r -s C:\boot.ini del C:\boot.ini BootCfg /Rebuild I added /fastdetect when it asked for options. I re-ran Windows setup - no change - no repair option. UPDATE: I followed the procedure at http://support.microsoft.com/default.aspx?scid=kb;en-us;307545 I rebooted. I now get a quick message on bootup to select the boot - 1: [blank] ; Windows XP Professional ; Windows Recover Console. The "1: " is new. The rest is the way it was when all was okay. Selecting 1: and the next one gives the same result - I get to a login icon, and then it asks for a password, with the blinking cursor, but I can't type anything. I reboot with the Windows CD. Now I see a repair option for installation "1: " I selected R on that, and it did "Setup is copying files..." and rebooted when it was done. Then it booted, and I got a window saying "Setup will complete in approximately 39 minutes." That's where I am now. I wasn't expecting this last part - I did a repair several months ago and I don't recall that. UPDATE: Booted up. Asked if I wanted to register Windows online. All my icons are there, and the old desktop documents. Good. All the applications I tried from the Start Menu work (tested a few), except Corel Photopaint - I get registry entry not found errors. Windows ran for a while, then froze. The mouse and keyboard don't work. Pressing the power button got Windows to shut down. I probably need to put SP2 on it, and then all the updates for my laptop for XP Pro SP2 (drivers), there's a bunch. The mouse and keyboard quit working again. That wasn't a problem when I first set up this laptop. I've ran 4 times now. Two mouse/keyboards hangs by pressing Ctrl-C (to copy text from a notepad document), and two by selecting Start-Run (wasn't able to type anything in the box).

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Could not find rake-10.1.0 in any of the sources

- by spuder

I've got a ruby on rails application (gitlab) which is installed via puppet. Everything on the test system runs fine, but production generates an error about rake Running /home/git/gitlab-shell/bin/check Could not find rake-10.1.0 in any of the sources Run bundle install to install missing gems. Here is the full rake check: root@gitlab:/home/git# sudo -u git -H bundle exec rake gitlab:check RAILS_ENV=production Checking Environment ... Git configured for git user? ... yes Has python2? ... yes python2 is supported version? ... yes Checking Environment ... Finished Checking GitLab Shell ... GitLab Shell version >= 1.7.1 ? ... OK (1.7.1) Repo base directory exists? ... yes Repo base directory is a symlink? ... no Repo base owned by git:git? ... yes Repo base access is drwxrws---? ... yes update hook up-to-date? ... yes update hooks in repos are links: ... Could not find rake-10.1.0 in any of the sources Run `bundle install` to install missing gems. gitlab-shell self-check failed Try fixing it: Make sure GitLab is running; Check the gitlab-shell configuration file: sudo -u git -H editor /home/git/gitlab-shell/config.yml Please fix the error above and rerun the checks. Checking GitLab Shell ... Finished Checking Sidekiq ... Running? ... yes Number of Sidekiq processes ... 1 Checking Sidekiq ... Finished Checking GitLab ... Database config exists? ... yes Database is SQLite ... no All migrations up? ... yes GitLab config exists? ... yes GitLab config outdated? ... no Log directory writable? ... yes Tmp directory writable? ... yes Init script exists? ... yes Init script up-to-date? ... yes projects have namespace: ... Spencer Owen / bar ... yes Projects have satellites? ... Spencer Owen / bar ... can't create, repository is empty Redis version >= 2.0.0? ... yes Your git bin path is "/usr/bin/git" Git version >= 1.7.10 ? ... yes (1.8.4) Checking GitLab ... Finished The step 'gitlab-shell check' effectively runs the following command. If I run that command manually, everything passes. root@gitlab:/home/git/gitlab# sudo -u git -H /home/git/gitlab-shell/bin/check Check GitLab API access: OK Check directories and files: /home/git/repositories: OK /home/git/.ssh/authorized_keys: OK I have verified that rake is in fact installed root@gitlab:/home/git/gitlab# gem install rake -v 10.1.0 root@gitlab:/home/git/gitlab# bundle install root@gitlab:/home/git/gitlab# sudo -u git -H gem install rake -v 10.1.0 root@gitlab:/home/git/gitlab# sudo -u git -H bundle install Ruby is installed with update alternatives root@gitlab:/home/git/gitlab# sudo -u git -H ruby --version ruby 1.9.3p0 (2011-10-30 revision 33570) [x86_64-linux] root@gitlab:/home/git/gitlab# sudo -u git -H ls -l `which ruby` lrwxrwxrwx 1 root root 22 Oct 8 20:26 /usr/bin/ruby -> /etc/alternatives/ruby root@gitlab:/home/git/gitlab# sudo -u git -H gem --version 2.1.10 root@gitlab:/home/git/gitlab# sudo -u git -H ls -l `which gem` lrwxrwxrwx 1 root root 21 Oct 10 20:50 /usr/bin/gem -> /etc/alternatives/gem I've tried the solution mentioned below, to allow shared gems http://stackoverflow.com/questions/19284914/bundle-exec-fails-with-could-not-find-rake-10-1-0-in-any-of-the-sources http://stackoverflow.com/questions/18978002/could-not-find-rake-with-bundle-exec root@gitlab:/home/git/gitlab# cat /home/git/gitlab/.bundle/config --- BUNDLE_FROZEN: '1' BUNDLE_PATH: vendor/bundle BUNDLE_WITHOUT: development:test:postgres BUNDLE_DISABLE_SHARED_GEMS: '1' I've exhausted google, so I'm hoping for someone more familiar with ruby to offer any ideas how to resolve the error. Could not find rake-10.1.0 in any of the sources

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Active Directory: trouble adding new DC

- by ethrbunny

I have a domain with 3 DCs. One is starting to fail so I brought up a new one. All are running Win 2003. Problem: there appear to be replication issues between the 4 machines but I can't figure out what's causing this. All are registered with the DNS as identically as I can make them. How do I know there is a problem? Nagios is telling me that the other 3 DCs are having KCCEvent errors and the new machine is reporting "failed connectivity" errors. Doing dcdiag on the new machine reports: the host could not be resolved to an IP address. This seems crazy as I log into it using the DNS name. I can ping it from the other three machines using this DNS name as well. repadmin /showreps from the new machine says its seeing the other 3 machines. Doing the same from one of the older machines doesn't show the new machine. I've tried netdiag /repair numerous times. No luck. There are no firewalls running on any of the machines. If I look at Domain info via MMC (on the new machine) it appears that all the information is current. Users, computers, DCs.. its all there. Im puzzled as to what step(s) I've missed in adding this new machine. Suggestions? EDIT: dcdiag from non-working: C:\Documents and Settings\Administrator.BME>dcdiag Domain Controller Diagnosis Performing initial setup: Done gathering initial info. Doing initial required tests Testing server: Default-First-Site-Name\YELLOW Starting test: Connectivity The host 312ce6ea-7909-4e15-aff6-45c3d1d9a0d9._msdcs.server.edu could not be resolved to an IP address. Check the DNS server, DHCP, server name, etc Although the Guid DNS name (312ce6ea-7909-4e15-aff6-45c3d1d9a0d9._msdcs.server.edu) couldn't be resolved, the server name (yellow.server.edu) resolved to the IP address (10.127.24.79) and was pingable. Check that the IP address is registered correctly with the DNS server. ......................... YELLOW failed test Connectivity Doing primary tests Testing server: Default-First-Site-Name\YELLOW Skipping all tests, because server YELLOW is not responding to directory service requests Running partition tests on : Schema Starting test: CrossRefValidation ......................... Schema passed test CrossRefValidation Starting test: CheckSDRefDom ......................... Schema passed test CheckSDRefDom Running partition tests on : Configuration Starting test: CrossRefValidation ......................... Configuration passed test CrossRefValidation Starting test: CheckSDRefDom ......................... Configuration passed test CheckSDRefDom Running partition tests on : bme Starting test: CrossRefValidation ......................... bme passed test CrossRefValidation Starting test: CheckSDRefDom ......................... bme passed test CheckSDRefDom Running enterprise tests on : server.edu Starting test: Intersite ......................... server.edu passed test Intersite Starting test: FsmoCheck ......................... server.edu passed test FsmoCheck dcdiag from working: P:\>dcdiag Domain Controller Diagnosis Performing initial setup: Done gathering initial info. Doing initial required tests Testing server: Default-First-Site-Name\AD1 Starting test: Connectivity ......................... AD1 passed test Connectivity Doing primary tests Testing server: Default-First-Site-Name\AD1 Starting test: Replications ......................... AD1 passed test Replications Starting test: NCSecDesc ......................... AD1 passed test NCSecDesc Starting test: NetLogons ......................... AD1 passed test NetLogons Starting test: Advertising ......................... AD1 passed test Advertising Starting test: KnowsOfRoleHolders ......................... AD1 passed test KnowsOfRoleHolders Starting test: RidManager ......................... AD1 passed test RidManager Starting test: MachineAccount ......................... AD1 passed test MachineAccount Starting test: Services ......................... AD1 passed test Services Starting test: ObjectsReplicated ......................... AD1 passed test ObjectsReplicated Starting test: frssysvol ......................... AD1 passed test frssysvol Starting test: frsevent ......................... AD1 passed test frsevent Starting test: kccevent ......................... AD1 passed test kccevent Starting test: systemlog ......................... AD1 passed test systemlog Starting test: VerifyReferences ......................... AD1 passed test VerifyReferences Running partition tests on : Schema Starting test: CrossRefValidation ......................... Schema passed test CrossRefValidation Starting test: CheckSDRefDom ......................... Schema passed test CheckSDRefDom Running partition tests on : Configuration Starting test: CrossRefValidation ......................... Configuration passed test CrossRefValidation Starting test: CheckSDRefDom ......................... Configuration passed test CheckSDRefDom Running partition tests on : bme Starting test: CrossRefValidation ......................... bme passed test CrossRefValidation Starting test: CheckSDRefDom ......................... bme passed test CheckSDRefDom Running enterprise tests on : server.edu Starting test: Intersite ......................... server.edu passed test Intersite Starting test: FsmoCheck ......................... server.edu passed test FsmoCheck P:\>

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dual boot install--no GRUB

- by Jim Syyap

My computer recently had a hardware upgrade and now runs on Windows 7. I decided to install Ubuntu 11.04 as dual boot using the ISO I got from ubuntu.com downloaded onto my USB stick. Restarting with the USB stick, I was able to install Ubuntu 11.04 choosing the option: Install Ubuntu 11.04 side by side with Windows 7 (or something like that). No errors were encountered on installation. However on restarting, there was no GRUB; the system went straight into Windows 7. Looking for answers, I found these: http://essayboard.com/2011/07/12/how-to-dual-boot-ubuntu-11-04-and-windows-7-the-traditional-way-through-grub-2/ http://ubuntuforums.org/showthread.php?t=1774523 Following their instructions, I got: Boot Info Script 0.60 from 17 May 2011 ============================= Boot Info Summary: =============================== => Windows is installed in the MBR of /dev/sda. => Syslinux MBR (3.61-4.03) is installed in the MBR of /dev/sdb. => Grub2 (v1.99) is installed in the MBR of /dev/sdc and looks at sector 1 of the same hard drive for core.img. core.img is at this location and looks for (,msdos7)/boot/grub on this drive. sda1: __________________________________________________ ________________________ File system: ntfs Boot sector type: Windows Vista/7 Boot sector info: No errors found in the Boot Parameter Block. Operating System: Boot files: /grldr /bootmgr /Boot/BCD /grldr sda2: __________________________________________________ ________________________ File system: ntfs Boot sector type: Windows Vista/7 Boot sector info: No errors found in the Boot Parameter Block. Operating System: Windows 7 Boot files: /Windows/System32/winload.exe sdb1: __________________________________________________ ________________________ File system: vfat Boot sector type: SYSLINUX 4.02 debian-20101016 ...........>...r>....... ......0...~.k...~...f...M.f.f....f..8~....>2} Boot sector info: Syslinux looks at sector 1437504 of /dev/sdb1 for its second stage. SYSLINUX is installed in the directory. The integrity check of the ADV area failed. According to the info in the boot sector, sdb1 starts at sector 0. But according to the info from fdisk, sdb1 starts at sector 62. Operating System: Boot files: /boot/grub/grub.cfg /syslinux/syslinux.cfg /ldlinux.sys sdc1: __________________________________________________ ________________________ File system: ntfs Boot sector type: Windows XP Boot sector info: No errors found in the Boot Parameter Block. Operating System: Boot files: sdc2: __________________________________________________ ________________________ File system: Extended Partition Boot sector type: - Boot sector info: sdc5: __________________________________________________ ________________________ File system: swap Boot sector type: - Boot sector info: sdc6: __________________________________________________ ________________________ File system: swap Boot sector type: - Boot sector info: sdc7: __________________________________________________ ________________________ File system: ext4 Boot sector type: - Boot sector info: Operating System: Ubuntu 11.04 Boot files: /boot/grub/grub.cfg /etc/fstab /boot/grub/core.img sdc8: __________________________________________________ ________________________ File system: swap Boot sector type: - Boot sector info: Going back into Ubuntu and running sudo fdisk -l , I got these: ubuntu@ubuntu:~$ sudo fdisk -l Disk /dev/sda: 160.0 GB, 160041885696 bytes 255 heads, 63 sectors/track, 19457 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x0002f393 Device Boot Start End Blocks Id System /dev/sda1 * 1 13 102400 7 HPFS/NTFS Partition 1 does not end on cylinder boundary. /dev/sda2 13 19458 156185600 7 HPFS/NTFS Disk /dev/sdb: 2011 MB, 2011168768 bytes 62 heads, 62 sectors/track, 1021 cylinders Units = cylinders of 3844 * 512 = 1968128 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x000f2ab9 Device Boot Start End Blocks Id System /dev/sdb1 * 1 1021 1962331 c W95 FAT32 (LBA) Disk /dev/sdc: 1000.2 GB, 1000202043392 bytes 255 heads, 63 sectors/track, 121600 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00261ddd Device Boot Start End Blocks Id System /dev/sdc1 * 1 60657 487222656+ 7 HPFS/NTFS /dev/sdc2 60657 121600 489527681 5 Extended /dev/sdc5 120563 121600 8337703+ 82 Linux swap / Solaris /dev/sdc6 120073 120562 3930112 82 Linux swap / Solaris /dev/sdc7 60657 119584 473328640 83 Linux /dev/sdc8 119584 120072 3923968 82 Linux swap / Solaris Should I proceed and do the following? Assuming Ubuntu 11.04 was installed on device sdb1, do this: sudo mount /dev/sdb1 /mnt Then do this: sudo grub-install--root-directory=/mnt /dev/sdb Notice there are two dashes in front of the root directory, and I'm not using sdb1 but sdb. Since the command in step 15 had reinstalled Grub 2, now we need to unmount the /mnt (i.e. sdb1) to clean up. Do this: sudo umount /mnt Reboot and remove Ubuntu 11.04 CD/DVD from disk tray. Log into Ubuntu 11.04 (you have no choice but it will make you log into Ubuntu 11.04 at this point). Open up a terminal in Ubuntu 11.04 (using real installation, not live CD/DVD). Execute this command: sudo update-grub Reboot the machine.

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What needs updating when moving a bootable Windows 7 (or Vista) partition?

- by SuperTempel

When I move a bootable NTFS partition with Windows on it to a different block offset, what needs updating to make it bootable again? In particular, here's what I tried: I have a disk with several partitions, one of which is the NTFS partition with Windows on it, and the disk uses the plain old MBR block 0 for the partitions layout (no more than 4 partitions). Now I format and partition a new, larger, disk. There I make room for the NTFS partition and copy the contents from the old disk's NTFS Windows partition into. And I make the partition "active". However, when I try to boot from this disk, I get a "read error" message immediately and the booting stops, the exact text is: A disk read error occurred Press Ctrl+Alt+Del to restart I verified that both disks have the same boot sector code in block 0. It seems to me that something else might need updating. I guess that somewhere there's a absolute block reference that I need to update, probably pointing to the next level loader or to the NT kernel. Update: I found this article going quite into the depth of what I want to know. However, it says to modify boot.ini, but I have Windows 7 installed here, where such things appear to have changed: No boot.ini but a folder called System Volume Information with GUID and other data in it that sounds related to my problem. Going to keep digging... Update 2: Thanks to the terrible looking but very informative website by starman, I was able to figure out the first step: The NTFS boot sector has a field for "hidden" sectors. This feld has to contain the sector number of the boot sector. This solves the "read error" message. Now, however, I get a "BOOTMGR is missing" error instead. Looks like there's another place where a block number has to be adjusted, but I can't find anything in the code listing about this. I do find a lot of help sites suggesting Windows tools for fixing this "BOOTMGR is missing" problem, but none seem to know what goes on behind the scenes. Kind of like suggesting to re-install Windows when there's a little problem with it. At least, those fixes seem to work, mostly involving the Bcdedit and Bootrec tools. Now, who knows what they do, especially the latter, in regards to a moved partition? Update 3: After lots of trial-and-error attempts, I believe now that the solution lies in the BCD-Template registry file, residing usually inside \Windows\System32\config. If I get this updated using the "bcdboot" command, Windows starts up from it. I am now in the middle of figuring out what information this registry contains relevant to the above question. Any pointers to the contents of this registry are welcome. Update 4: Turns out that while the BCD-Template file gets rewritten and has different binary contents than its predecessor, the values inside do not change. So it must be something else that bcdboot.exe writes. I had previously already checked if it changes the first 32 boot blocks of the partition, but they appear to remain unchanged. Parititon map doesn't get changed, either. So what is it that bcdboot modifies besides the BCD registry? Any tips on how I can trace that? Are there low level tools that show me what files a program writes to? Update 5: The answer seems to be: c:\Boot\BCD is also changed, and that appears to be the key file for the boot manager's process. I'll investigate this later... Update 6: It seems to be an important detail that I had originally two partitions created when I installed Windows 7: A small partition of 204800 sectors which appears to be a bootstrap partition, followed by the actual, large, partition containing the Windows system (drive C:). When I tried to transfer this installation to a new, larger, disk, I had kept the same two partitions intact on the new drive, although they ended up at a different offset. This alone led to the "BOOTMGR is missing" message. Since then, I've used bcdboot.exe only on the Windows partition, which added the \Boot\BCD file on that partition. That file (and folder) did originally only exist on the smaller partition. Hence, this problem may be more complicated in my case as one partition (the boot strapper) referred to another partition (the one containing the OS), whereas other people may only have to deal with one partition containing both, and maybe there the solution is simpler. Update 7: Found one more detail: The \Boot\BCD file records the MBR's serial number. If that number doesn't match, the system won't boot. Next I'll test if there's also an absolute block reference stored in there.

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