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  • Why people don't patch and upgrade?!?

    - by Mike Dietrich
    Discussing the topic "Why Upgrade" or "Why not Upgrade" is not always fun. Actually the arguments repeat from customer to customer. Typically we hear things such as: A PSU or Patch Set introduces new bugs A new PSU or Patch Set introduces new features which lead to risk and require application verification  Patching means risk Patching changes the execution plans Patching requires too much testing Patching is too much work for our DBAs Patching costs a lot of money and doesn't pay out And to be very honest sometimes it's hard for me to stay calm in such discussions. Let's discuss some of these points a bit more in detail. A PSU or Patch Set introduces new bugsWell, yes, that is true as no software containing more than some lines of code is bug free. This applies to Oracle's code as well as too any application or operating system code. But first of all, does that mean you never patch your OS because the patch may introduce new flaws? And second, what is the point of saying "it introduces new bugs"? Does that mean you will never get rid of the mean issues we know about and we fixed already? Scroll down from MOS Note:161818.1 to the patch release you are on, no matter if it's 10.2.0.4 or 11.2.0.3 and check for the Known Issues And Alerts.Will you take responsibility to know about all these issues and refuse to upgrade to 11.2.0.4? I won't. A new PSU or Patch Set introduces new featuresOk, we can discuss that. Offering new functionality within a database patch set is a dubious thing. It has advantages such as in 11.2.0.4 where we backported Database Redaction to. But this is something you will only use once you have an Advanced Security license. I interpret that statement I've heard quite often from customers in a different way: People don't want to get surprises such as new behaviour. This certainly gives everybody a hard time. And we've had many examples in the past (SESSION_CACHED_CURSROS in 10.2.0.4,  _DATAFILE_WRITE_ERRORS_CRASH_INSTANCE in 11.2.0.2 and others) where those things weren't documented, not even in the README. Thanks to many friends out there I learned about those as well. So new behaviour is the topic people consider as risky - not really new features. And just to point this out: A PSU never brings in new features or new behaviour by definition! Patching means riskDoes it really mean risk? Yes, there were issues in the past (and sometimes in the present as well) where a patch didn't get installed correctly. But personally I consider it way more risky to not patch. Keep that in mind: The day Oracle publishes an PSU (or CPU) containing security fixes all the great security experts out there go public with their findings as well. So from that day on even my grandma can find out about those issues and try to attack somebody. Now a lot of people say: "My database does not face the internet." And I will answer: "The enemy is sitting already behind your firewalls. And knows potentially about these things." My statement: Not patching introduces way more risk to your environment than patching. Seriously! Patching changes the execution plansDo they really? I agree - there's a very small risk for this happening with Patch Sets. But not with PSUs or CPUs as they contain no optimizer fixes changing behaviour (but they may contain fixes curing wrong-query-result-bugs). But what's the point of a changing execution plan? In Oracle Database 11g it is so simple to be prepared. SQL Plan Management is a free EE feature - so once that occurs you'll put the plan into the Plan Baseline. Basta! Yes, you wouldn't like to get such surprises? Than please use the SQL Performance Analyzer (SPA) from Real Application Testing and you'll detect that easily upfront in minutes. And not to forget this, a plan change can also be very positive!Yes, there's a little risk with a database patchset - and we have many possibilites to detect this before patching. Patching requires too much testingWell, does it really? I have seen in the past 12 years how people test. There are very different efforts and approaches on this. I have seen people spending a hell of money on licenses or on project team staffing. And I have seen people sailing blindly without any tests just going the John-Wayne-approach.Proper tools will allow you to test easily without too much efforts. See the paragraph above. We have used Real Application Testing in so many customer projects reducing the amount of work spend on testing by over 50%. But apart from that at some point you will have to stop testing. If you don't you'll get lost and you'll burn money. There's no 100% guaranty. You will have to deal with a little risk as reaching the final 5% of certainty will cost you the same as it did cost to reach 95%. And doing this will lead to abnormal long product cycles that you'll run behind forever. And this will cost even more money. Patching is too much work for our DBAsPatching is a lot of work. I agree. And it's no fun work. It's boring, annoying. You don't learn much from that. That's why you should try to automate this task. Use the Database's Lifecycle Management Pack. And don't cry about the fact that it costs money. Yes it does. But it will ease the process and you'll save a lot of costs as you don't waste your valuable time with patching. Or use Oracle Database 12c Oracle Multitenant and patch either by unplug/plug or patch an entire container database with all PDBs with one patch in one task. We have customer reference cases proofing it saved them 75% of time, effort and cost since they've used Lifecycle Management Pack. So why don't you use it? Patching costs a lot of money and doesn't pay outWell, see my statements in the paragraph above. And it pays out as flying with a database with 100 known critical flaws in it which are already fixed by Oracle (such as in the Oct 2013 PSU for Oracle Database 12c) will cost ways more in case of failure or even data loss. Bet with me? Let me finally ask you some questions. What cell phone are you using and which OS does it run? Do you have an iPhone 5 and did you upgrade already to iOS 7.0.3? I've just encountered on mine that the alarm (which I rely on when traveling) has gotten now a dependency on the physical switch "sound on/off". If it is switched to "off" physically the alarm rings "silently". What a wonderful example of a behaviour change coming in with a patch set. Will this push you to stay with iOS5 or iOS6? No, because those have security flaws which won't be fixed anymore. What browser are you surfing with? Do you use Mozilla 3.6? Well, congratulations to all the hackers. It will be easy for them to attack you and harm your system. I'd guess you have the auto updater on.  Same for Google Chrome, Safari, IE. Right? -Mike The T.htmtableborders, .htmtableborders td, .htmtableborders th {border : 1px dashed lightgrey ! important;} html, body { border: 0px; } body { background-color: #ffffff; } img, hr { cursor: default }

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  • Apache on Win32: Slow Transfers of single, static files in HTTP, fast in HTTPS

    - by Michael Lackner
    I have a weird problem with Apache 2.2.15 on Windows 2000 Server SP4. Basically, I am trying to serve larger static files, images, videos etc. The download seems to be capped at around 550kB/s even over 100Mbit LAN. I tried other protocols (FTP/FTPS/FTP+ES/SCP/SMB), and they are all in the multi-megabyte range. The strangest thing is that, when using Apache with HTTPS instead of HTTP, it serves very fast, around 2.7MByte/s! I also tried the AnalogX SimpleWWW server just to test the plain HTTP speed of it, and it gave me a healthy 3.3Mbyte/s. I am at a total loss here. I searched the web, and tried to change the following Apache configuration directives in httpd.conf, one at a time, mostly to no avail at all: SendBufferSize 1048576 #(tried multiples of that too, up to 100Mbytes) EnableSendfile Off #(minor performance boost) EnableMMAP Off Win32DisableAcceptEx HostnameLookups Off #(default) I also tried to tune the following registry parameters, setting their values to 4194304 in decimal (they are REG_DWORD), and rebooting afterwards: HKLM\SYSTEM\CurrentControlSet\Services\AFD\Parameters\DefaultReceiveWindow HKLM\SYSTEM\CurrentControlSet\Services\AFD\Parameters\DefaultSendWindow Additionally, I tried to install mod_bw, which sets the event timer precision to 1ms, and allows for bandwidth throttling. According to some people it boosts static file serving performance when set to unlimited bandwidth for everybody. Unfortunately, it did nothing for me. So: AnalogX HTTP: 3300kB/s Gene6 FTPD, plain: 3500kB/s Gene6 FTPD, Implicit and Explicit SSL, AES256 Cipher: 1800-2000kB/s freeSSHD: 1100kB/s SMB shared folder: about 3000kB/s Apache HTTP, plain: 550kB/s Apache HTTPS: 2700kB/s Clients that were used in the bandwidth testing: Internet Explorer 8 (HTTP, HTTPS) Firefox 8 (HTTP, HTTPS) Chrome 13 (HTTP, HTTPS) Opera 11.60 (HTTP, HTTPS) wget under CygWin (HTTP, HTTPS) FileZilla (FTP, FTPS, FTP+ES, SFTP) Windows Explorer (SMB) Generally, transfer speeds are not too high, but that's because the server machine is an old quad Pentium Pro 200MHz machine with 2GB RAM. However, I would like Apache to serve at at least 2Mbyte/s instead of 550kB/s, and that already works with HTTPS easily, so I fail to see why plain HTTP is so crippled. I am using a Kerio Winroute Firewall, but no Throttling and no special filters peeking into HTTP traffic, just the plain Firewall functionality for blocking/allowing connections. The Apache error.log (Loglevel info) shows no warnings, no errors. Also nothing strange to be seen in access.log. I have already stripped down my httpd.conf to the bare minimum just to make sure nothing is interfering, but that didn't help either. If you have any idea, help would be greatly appreciated, since I am totally out of ideas! Thanks! Edit: I have now tried a newer Apache 2.2.21 to see if it makes any difference. However, the behaviour is exactly the same. Edit 2: KM01 has requested a sniff on the HTTP headers, so here comes the LiveHTTPHeaders output (an extension to Firefox). The Output is generated on downloading a single file called "elephantsdream_source.264", which is an H.264/AVC elementary video stream under an Open Source license. I have taken the freedom to edit the URL, removing folders and changing the actual servers domain name to www.mydomain.com. Here it is: LiveHTTPHeaders, Plain HTTP: http://www.mydomain.com/elephantsdream_source.264 GET /elephantsdream_source.264 HTTP/1.1 Host: www.mydomain.com User-Agent: Mozilla/5.0 (Windows NT 5.2; WOW64; rv:6.0.2) Gecko/20100101 Firefox/6.0.2 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Accept-Language: de-de,de;q=0.8,en-us;q=0.5,en;q=0.3 Accept-Encoding: gzip, deflate Accept-Charset: ISO-8859-1,utf-8;q=0.7,*;q=0.7 Connection: keep-alive HTTP/1.1 200 OK Date: Wed, 21 Dec 2011 20:55:16 GMT Server: Apache/2.2.21 (Win32) mod_ssl/2.2.21 OpenSSL/0.9.8r PHP/5.2.17 Last-Modified: Thu, 28 Oct 2010 20:20:09 GMT Etag: "c000000013fa5-29cf10e9-493b311889d3c" Accept-Ranges: bytes Content-Length: 701436137 Keep-Alive: timeout=15, max=100 Connection: Keep-Alive Content-Type: text/plain LiveHTTPHeaders, HTTPS: https://www.mydomain.com/elephantsdream_source.264 GET /elephantsdream_source.264 HTTP/1.1 Host: www.mydomain.com User-Agent: Mozilla/5.0 (Windows NT 5.2; WOW64; rv:6.0.2) Gecko/20100101 Firefox/6.0.2 Accept: text/html,application/xhtml+xml,application/xml;q=0.9,*/*;q=0.8 Accept-Language: de-de,de;q=0.8,en-us;q=0.5,en;q=0.3 Accept-Encoding: gzip, deflate Accept-Charset: ISO-8859-1,utf-8;q=0.7,*;q=0.7 Connection: keep-alive HTTP/1.1 200 OK Date: Wed, 21 Dec 2011 20:56:57 GMT Server: Apache/2.2.21 (Win32) mod_ssl/2.2.21 OpenSSL/0.9.8r PHP/5.2.17 Last-Modified: Thu, 28 Oct 2010 20:20:09 GMT Etag: "c000000013fa5-29cf10e9-493b311889d3c" Accept-Ranges: bytes Content-Length: 701436137 Keep-Alive: timeout=15, max=100 Connection: Keep-Alive Content-Type: text/plain

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  • Why do we (really) program to interfaces?

    - by Kyle Burns
    One of the earliest lessons I was taught in Enterprise development was "always program against an interface".  This was back in the VB6 days and I quickly learned that no code would be allowed to move to the QA server unless my business objects and data access objects each are defined as an interface and have a matching implementation class.  Why?  "It's more reusable" was one answer.  "It doesn't tie you to a specific implementation" a slightly more knowing answer.  And let's not forget the discussion ending "it's a standard".  The problem with these responses was that senior people didn't really understand the reason we were doing the things we were doing and because of that, we were entirely unable to realize the intent behind the practice - we simply used interfaces and had a bunch of extra code to maintain to show for it. It wasn't until a few years later that I finally heard the term "Inversion of Control".  Simply put, "Inversion of Control" takes the creation of objects that used to be within the control (and therefore a responsibility of) of your component and moves it to some outside force.  For example, consider the following code which follows the old "always program against an interface" rule in the manner of many corporate development shops: 1: ICatalog catalog = new Catalog(); 2: Category[] categories = catalog.GetCategories(); In this example, I met the requirement of the rule by declaring the variable as ICatalog, but I didn't hit "it doesn't tie you to a specific implementation" because I explicitly created an instance of the concrete Catalog object.  If I want to test the functionality of the code I just wrote I have to have an environment in which Catalog can be created along with any of the resources upon which it depends (e.g. configuration files, database connections, etc) in order to test my functionality.  That's a lot of setup work and one of the things that I think ultimately discourages real buy-in of unit testing in many development shops. So how do I test my code without needing Catalog to work?  A very primitive approach I've seen is to change the line the instantiates catalog to read: 1: ICatalog catalog = new FakeCatalog();   once the test is run and passes, the code is switched back to the real thing.  This obviously poses a huge risk for introducing test code into production and in my opinion is worse than just keeping the dependency and its associated setup work.  Another popular approach is to make use of Factory methods which use an object whose "job" is to know how to obtain a valid instance of the object.  Using this approach, the code may look something like this: 1: ICatalog catalog = CatalogFactory.GetCatalog();   The code inside the factory is responsible for deciding "what kind" of catalog is needed.  This is a far better approach than the previous one, but it does make projects grow considerably because now in addition to the interface, the real implementation, and the fake implementation(s) for testing you have added a minimum of one factory (or at least a factory method) for each of your interfaces.  Once again, developers say "that's too complicated and has me writing a bunch of useless code" and quietly slip back into just creating a new Catalog and chalking any test failures up to "it will probably work on the server". This is where software intended specifically to facilitate Inversion of Control comes into play.  There are many libraries that take on the Inversion of Control responsibilities in .Net and most of them have many pros and cons.  From this point forward I'll discuss concepts from the standpoint of the Unity framework produced by Microsoft's Patterns and Practices team.  I'm primarily focusing on this library because it questions about it inspired this posting. At Unity's core and that of most any IoC framework is a catalog or registry of components.  This registry can be configured either through code or using the application's configuration file and in the most simple terms says "interface X maps to concrete implementation Y".  It can get much more complicated, but I want to keep things at the "what does it do" level instead of "how does it do it".  The object that exposes most of the Unity functionality is the UnityContainer.  This object exposes methods to configure the catalog as well as the Resolve<T> method which is used to obtain an instance of the type represented by T.  When using the Resolve<T> method, Unity does not necessarily have to just "new up" the requested object, but also can track dependencies of that object and ensure that the entire dependency chain is satisfied. There are three basic ways that I have seen Unity used within projects.  Those are through classes directly using the Unity container, classes requiring injection of dependencies, and classes making use of the Service Locator pattern. The first usage of Unity is when classes are aware of the Unity container and directly call its Resolve method whenever they need the services advertised by an interface.  The up side of this approach is that IoC is utilized, but the down side is that every class has to be aware that Unity is being used and tied directly to that implementation. Many developers don't like the idea of as close a tie to specific IoC implementation as is represented by using Unity within all of your classes and for the most part I agree that this isn't a good idea.  As an alternative, classes can be designed for Dependency Injection.  Dependency Injection is where a force outside the class itself manipulates the object to provide implementations of the interfaces that the class needs to interact with the outside world.  This is typically done either through constructor injection where the object has a constructor that accepts an instance of each interface it requires or through property setters accepting the service providers.  When using dependency, I lean toward the use of constructor injection because I view the constructor as being a much better way to "discover" what is required for the instance to be ready for use.  During resolution, Unity looks for an injection constructor and will attempt to resolve instances of each interface required by the constructor, throwing an exception of unable to meet the advertised needs of the class.  The up side of this approach is that the needs of the class are very clearly advertised and the class is unaware of which IoC container (if any) is being used.  The down side of this approach is that you're required to maintain the objects passed to the constructor as instance variables throughout the life of your object and that objects which coordinate with many external services require a lot of additional constructor arguments (this gets ugly and may indicate a need for refactoring). The final way that I've seen and used Unity is to make use of the ServiceLocator pattern, of which the Patterns and Practices team has also provided a Unity-compatible implementation.  When using the ServiceLocator, your class calls ServiceLocator.Retrieve in places where it would have called Resolve on the Unity container.  Like using Unity directly, it does tie you directly to the ServiceLocator implementation and makes your code aware that dependency injection is taking place, but it does have the up side of giving you the freedom to swap out the underlying IoC container if necessary.  I'm not hugely concerned with hiding IoC entirely from the class (I view this as a "nice to have"), so the single biggest problem that I see with the ServiceLocator approach is that it provides no way to proactively advertise needs in the way that constructor injection does, allowing more opportunity for difficult to track runtime errors. This blog entry has not been intended in any way to be a definitive work on IoC, but rather as something to spur thought about why we program to interfaces and some ways to reach the intended value of the practice instead of having it just complicate your code.  I hope that it helps somebody begin or continue a journey away from being a "Cargo Cult Programmer".

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  • How I Work: Staying Productive Whilst Traveling

    - by BuckWoody
    I travel a lot. Not like some folks that are gone every week, mind you, although in the last month I’ve been to: Cambridge, UK; Anchorage, AK; San Jose, CA; Copenhagen, DK, Boston, MA; and I’m currently en-route to Anaheim, CA.  While this many places in a month is a bit unusual for me, I would say I travel frequently. I’ve travelled most of my 28+ years in IT, and at one time was a consultant traveling weekly.   With that much time away from my primary work location, I have to find ways to stay productive. Some might say “just rest – take a nap!” – but I’m not able to do that. For one thing, I’m a very light sleeper and I’ve never slept on a plane - even a 30+ hour trip to New Zealand in Business Class - so that just isn’t option. I also am not always in the plane, of course. There’s the hotel, the taxi/bus/train, the airport and then all that over again when I arrive. Since my regular jobs have many demands, I have to get work done.   Note: No, I’m not always focused on work. I need downtime just like everyone else. Sometimes I just think, watch a movie or listen to tunes – and I give myself permission to do that anytime – sometimes the whole trip. I have too fewheartbeats left in life to only focus on work – it’s just not that important, and neither am I. Some of these tasks are letters to friends and family, or other personal things. What I’m talking about here is a plan, not some task list I have to follow. When I get to the location I’m traveling to, I always build in as much time as I can to ensure I enjoy those sights and the people I’m with. I would find traveling to be a waste if not for that.   The Unrealistic Expectation As I would evaluate the trip I was taking – say a 6-8 hour flight – I would expect to get 10-12 hours of work done. After all, there’s the time at the airport, the taxi and so on, and then of course the time in the air with all of the room, power, internet and everything else I needed to get my work done. I would pile up tasks at home, pack my bags, and head happily to the magical land of the TSA.   Right. On return from the trip, I had accomplished little, had more e-mails and other work that had piled up, and I was tired, hungry, and unorganized. This had to change. So, I decided to do three things: Segment my work Set realistic expectations Plan accordingly  Segmenting By Available Resources The first task was to decide what kind of work I could do in each location – if any. I found that I was dependent on a few things to get work done, such as power, the Internet, and a place to sit down. Before I fly, I take some time at home to get all of the work I’d like to accomplish while away segmented into these areas, and print that out on paper, which goes in my suit-coat pocket along with a mechanical pencil. I print my tickets, and I’m all set for the adventure ahead. Then I simply do each kind of work whenever I’m in that situation. No power There are certain times when I don’t have power available. But not only that, I might not even be able to use most of my electronics. So I now schedule as many phone calls as I can for the taxi/bus/train ride and the airports as I can. I have a paper notebook (Moleskine, of course) and a pencil and I print out any notes or numbers I need prior to the trip. Once I’m airborne or at the airport, I work on my laptop. I check and respond to e-mails, create slides, write code, do architecture, whatever I can.  If I can’t use any electronics, or once the power runs out, I schedule time for reading. I can read at the airport or anywhere, actually, even in-flight or any other transport. I “read with a pencil”, meaning I take a lot of notes, which I liketo put in OneNote, but since in most cases I don’t have power, I use the Moleskine to do that. Speaking of which, sometimes as I’m thinking I come up with new topics, ideas, blog posts, or things to teach in my classes. Once again I take out the notebook and write it down. All of these notes get a check-mark when I get back to the office and transfer the writing to OneNote. I’ve tried those “smart pens” and so on to automate this, but it just never works out. Pencil and paper are just fine. As I mentioned, sometime I just need to think. I’ll do nothing, and let my mind wander, thinking of nothing in particular, or some math problem or science question I’m interested in. My only issue with this is that I communicate tothink, and I don’t want to drive people crazy by being that guy that won’t shut up, so I think in a different way. Power, but no Internet or Phone If I have power but no Internet or phone, I focus on the laptop and the tablet as before, and I also recharge my other gadgets. Power, Internet, Phone and a Place to Work At first I thought that when I arrived at the hotel or event I could get the same amount of work done that I do at the office. Not so. There’s simply too many distractions, things you need, or other issues that allow this. Of course, Ican work on any device, read, think, write or whatever, but I am simply not as productive as I am in my home office. So I plan for about 25-50% as much work getting done in this environment as I think I could really do. I’ve done some measurements, and this holds out to be true almost every time. The key is that I re-set my expectations (and my co-worker’s expectations as well) that this is the case. I use the Out-Of-Office notices to let people know that I’m just not going to be 100% at this time – it’s hard for everyone, but it’s more honest and realistic, and I’d rather they know that – and that I realize that – than to let them think I’m totally available. Because I’m not – I’m traveling. I don’t tend to put too much detail, because after all I don’t necessarily want to let people know when I’m not home :) but I do think it’s important to let people that depend on my know that I’ll get back with them later. I hope this helps you think through your own methodology of staying productive when you travel. Or perhaps you just go offline, and don’t worry about any of this – good for you! That’s completely valid as well.   (Oh, and yes, I wrote this at 35K feet, on Alaska Airlines on a trip. :)  Practice what you preach, Buck.)

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  • Disaster, or Migration?

    - by Rob Farley
    This post is in two parts – technical and personal. And I should point out that it’s prompted in part by this month’s T-SQL Tuesday, hosted by Allen Kinsel. First, the technical: I’ve had a few conversations with people recently about migration – moving a SQL Server database from one box to another (sometimes, but not primarily, involving an upgrade). One question that tends to come up is that of downtime. Obviously there will be some period of time between the old server being available and the new one. The way that most people seem to think of migration is this: Build a new server. Stop people from using the old server. Take a backup of the old server Restore it on the new server. Reconfigure the client applications (or alternatively, configure the new server to use the same address as the old) Make the new server online. There are other things involved, such as testing, of course. But this is essentially the process that people tell me they’re planning to follow. The bit that I want to look at today (as you’ve probably guessed from my title) is the “backup and restore” section. If a SQL database is using the Simple Recovery Model, then the only restore option is the last database backup. This backup could be full or differential. The transaction log never gets backed up in the Simple Recovery Model. Instead, it truncates regularly to stay small. One that’s using the Full Recovery Model (or Bulk-Logged) won’t truncate its log – the log must be backed up regularly. This provides the benefit of having a lot more option available for restores. It’s a requirement for most systems of High Availability, because if you’re making sure that a spare box is up-and-running, ready to take over, then you have to be interested in the logs that are happening on the current box, rather than truncating them all the time. A High Availability system such as Mirroring, Replication or Log Shipping will initialise the spare machine by restoring a full database backup (and maybe a differential backup if available), and then any subsequent log backups. Once the secondary copy is close, transactions can be applied to keep the two in sync. The main aspect of any High Availability system is to have a redundant system that is ready to take over. So the similarity for migration should be obvious. If you need to move a database from one box to another, then introducing a High Availability mechanism can help. By turning on the Full Recovery Model and then taking a backup (so that the now-interesting logs have some context), logs start being kept, and are therefore available for getting the new box ready (even if it’s an upgraded version). When the migration is ready to occur, a failover can be done, letting the new server take over the responsibility of the old, just as if a disaster had happened. Except that this is a planned failover, not a disaster at all. There’s a fine line between a disaster and a migration. Failovers can be useful in patching, upgrading, maintenance, and more. Hopefully, even an unexpected disaster can be seen as just another failover, and there can be an opportunity there – perhaps to get some work done on the principal server to increase robustness. And if I’ve just set up a High Availability system for even the simplest of databases, it’s not necessarily a bad thing. :) So now the personal: It’s been an interesting time recently... June has been somewhat odd. A court case with which I was involved got resolved (through mediation). I can’t go into details, but my lawyers tell me that I’m allowed to say how I feel about it. The answer is ‘lousy’. I don’t regret pursuing it as long as I did – but in the end I had to make a decision regarding the commerciality of letting it continue, and I’m going to look forward to the days when the kind of money I spent on my lawyers is small change. Mind you, if I had a similar situation with an employer, I’d do the same again, but that doesn’t really stop me feeling frustrated about it. The following day I had to fly to country Victoria to see my grandmother, who wasn’t expected to last the weekend. She’s still around a week later as I write this, but her 92-year-old body has basically given up on her. She’s been a Christian all her life, and is looking forward to eternity. We’ll all miss her though, and it’s hard to see my family grieving. Then on Tuesday, I was driving back to the airport with my family to come home, when something really bizarre happened. We were travelling down the freeway, just pulled out to go past a truck (farm-truck sized, not a semi-trailer), when a car-sized mass of metal fell off it. It was something like an industrial air-conditioner, but from where I was sitting, it was just a mass of spinning metal, like something out of a movie (one friend described it as “holidays by Michael Bay”). Somehow, and I’m really don’t know how, the part of it nearest us bounced high enough to clear the car, and there wasn’t even a scratch. We pulled over the check, and I was just thanking God that we’d changed lanes when we had, and that we remained unharmed. I had all kinds of thoughts about what could’ve happened if we’d had something that size land on the windscreen... All this has drilled home that while I feel that I haven’t provided as well for the family as I could’ve done (like by pursuing an expensive legal case), I shouldn’t even consider that I have proper control over things. I get to live life, and make decisions based on what I feel is right at the time. But I’m not going to get everything right, and there will be things that feel like disasters, some which could’ve been in my control and some which are very much beyond my control. The case feels like something I could’ve pursued differently, a disaster that could’ve been avoided in some way. Gran dying is lousy of course. An accident on the freeway would have been awful. I need to recognise that the worst disasters are ones that I can’t affect, and that I need to look at things in context – perhaps seeing everything that happens as a migration instead. Life is never the same from one day to the next. Every event has a before and an after – sometimes it’s clearly positive, sometimes it’s not. I remember good events in my life (such as my wedding), and bad (such as the loss of my father when I was ten, or the back injury I had eight years ago). I’m not suggesting that I know how to view everything from the “God works all things for good” perspective, but I am trying to look at last week as a migration of sorts. Those things are behind me now, and the future is in God’s hands. Hopefully I’ve learned things, and will be able to live accordingly. I’ve come through this time now, and even though I’ll miss Gran, I’ll see her again one day, and the future is bright.

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  • Solving Big Problems with Oracle R Enterprise, Part I

    - by dbayard
    Abstract: This blog post will show how we used Oracle R Enterprise to tackle a customer’s big calculation problem across a big data set. Overview: Databases are great for managing large amounts of data in a central place with rigorous enterprise-level controls.  R is great for doing advanced computations.  Sometimes you need to do advanced computations on large amounts of data, subject to rigorous enterprise-level concerns.  This blog post shows how Oracle R Enterprise enables R plus the Oracle Database enabled us to do some pretty sophisticated calculations across 1 million accounts (each with many detailed records) in minutes. The problem: A financial services customer of mine has a need to calculate the historical internal rate of return (IRR) for its customers’ portfolios.  This information is needed for customer statements and the online web application.  In the past, they had solved this with a home-grown application that pulled trade and account data out of their data warehouse and ran the calculations.  But this home-grown application was not able to do this fast enough, plus it was a challenge for them to write and maintain the code that did the IRR calculation. IRR – a problem that R is good at solving: Internal Rate of Return is an interesting calculation in that in most real-world scenarios it is impractical to calculate exactly.  Rather, IRR is a calculation where approximation techniques need to be used.  In this blog post, we will discuss calculating the “money weighted rate of return” but in the actual customer proof of concept we used R to calculate both money weighted rate of returns and time weighted rate of returns.  You can learn more about the money weighted rate of returns here: http://www.wikinvest.com/wiki/Money-weighted_return First Steps- Calculating IRR in R We will start with calculating the IRR in standalone/desktop R.  In our second post, we will show how to take this desktop R function, deploy it to an Oracle Database, and make it work at real-world scale.  The first step we did was to get some sample data.  For a historical IRR calculation, you have a balances and cash flows.  In our case, the customer provided us with several accounts worth of sample data in Microsoft Excel.      The above figure shows part of the spreadsheet of sample data.  The data provides balances and cash flows for a sample account (BMV=beginning market value. FLOW=cash flow in/out of account. EMV=ending market value). Once we had the sample spreadsheet, the next step we did was to read the Excel data into R.  This is something that R does well.  R offers multiple ways to work with spreadsheet data.  For instance, one could save the spreadsheet as a .csv file.  In our case, the customer provided a spreadsheet file containing multiple sheets where each sheet provided data for a different sample account.  To handle this easily, we took advantage of the RODBC package which allowed us to read the Excel data sheet-by-sheet without having to create individual .csv files.  We wrote ourselves a little helper function called getsheet() around the RODBC package.  Then we loaded all of the sample accounts into a data.frame called SimpleMWRRData. Writing the IRR function At this point, it was time to write the money weighted rate of return (MWRR) function itself.  The definition of MWRR is easily found on the internet or if you are old school you can look in an investment performance text book.  In the customer proof, we based our calculations off the ones defined in the The Handbook of Investment Performance: A User’s Guide by David Spaulding since this is the reference book used by the customer.  (One of the nice things we found during the course of this proof-of-concept is that by using R to write our IRR functions we could easily incorporate the specific variations and business rules of the customer into the calculation.) The key thing with calculating IRR is the need to solve a complex equation with a numerical approximation technique.  For IRR, you need to find the value of the rate of return (r) that sets the Net Present Value of all the flows in and out of the account to zero.  With R, we solve this by defining our NPV function: where bmv is the beginning market value, cf is a vector of cash flows, t is a vector of time (relative to the beginning), emv is the ending market value, and tend is the ending time. Since solving for r is a one-dimensional optimization problem, we decided to take advantage of R’s optimize method (http://stat.ethz.ch/R-manual/R-patched/library/stats/html/optimize.html). The optimize method can be used to find a minimum or maximum; to find the value of r where our npv function is closest to zero, we wrapped our npv function inside the abs function and asked optimize to find the minimum.  Here is an example of using optimize: where low and high are scalars that indicate the range to search for an answer.   To test this out, we need to set values for bmv, cf, t, emv, tend, low, and high.  We will set low and high to some reasonable defaults. For example, this account had a negative 2.2% money weighted rate of return. Enhancing and Packaging the IRR function With numerical approximation methods like optimize, sometimes you will not be able to find an answer with your initial set of inputs.  To account for this, our approach was to first try to find an answer for r within a narrow range, then if we did not find an answer, try calling optimize() again with a broader range.  See the R help page on optimize()  for more details about the search range and its algorithm. At this point, we can now write a simplified version of our MWRR function.  (Our real-world version is  more sophisticated in that it calculates rate of returns for 5 different time periods [since inception, last quarter, year-to-date, last year, year before last year] in a single invocation.  In our actual customer proof, we also defined time-weighted rate of return calculations.  The beauty of R is that it was very easy to add these enhancements and additional calculations to our IRR package.)To simplify code deployment, we then created a new package of our IRR functions and sample data.  For this blog post, we only need to include our SimpleMWRR function and our SimpleMWRRData sample data.  We created the shell of the package by calling: To turn this package skeleton into something usable, at a minimum you need to edit the SimpleMWRR.Rd and SimpleMWRRData.Rd files in the \man subdirectory.  In those files, you need to at least provide a value for the “title” section. Once that is done, you can change directory to the IRR directory and type at the command-line: The myIRR package for this blog post (which has both SimpleMWRR source and SimpleMWRRData sample data) is downloadable from here: myIRR package Testing the myIRR package Here is an example of testing our IRR function once it was converted to an installable package: Calculating IRR for All the Accounts So far, we have shown how to calculate IRR for a single account.  The real-world issue is how do you calculate IRR for all of the accounts?This is the kind of situation where we can leverage the “Split-Apply-Combine” approach (see http://www.cscs.umich.edu/~crshalizi/weblog/815.html).  Given that our sample data can fit in memory, one easy approach is to use R’s “by” function.  (Other approaches to Split-Apply-Combine such as plyr can also be used.  See http://4dpiecharts.com/2011/12/16/a-quick-primer-on-split-apply-combine-problems/). Here is an example showing the use of “by” to calculate the money weighted rate of return for each account in our sample data set.  Recap and Next Steps At this point, you’ve seen the power of R being used to calculate IRR.  There were several good things: R could easily work with the spreadsheets of sample data we were given R’s optimize() function provided a nice way to solve for IRR- it was both fast and allowed us to avoid having to code our own iterative approximation algorithm R was a convenient language to express the customer-specific variations, business-rules, and exceptions that often occur in real-world calculations- these could be easily added to our IRR functions The Split-Apply-Combine technique can be used to perform calculations of IRR for multiple accounts at once. However, there are several challenges yet to be conquered at this point in our story: The actual data that needs to be used lives in a database, not in a spreadsheet The actual data is much, much bigger- too big to fit into the normal R memory space and too big to want to move across the network The overall process needs to run fast- much faster than a single processor The actual data needs to be kept secured- another reason to not want to move it from the database and across the network And the process of calculating the IRR needs to be integrated together with other database ETL activities, so that IRR’s can be calculated as part of the data warehouse refresh processes In our next blog post in this series, we will show you how Oracle R Enterprise solved these challenges.

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  • Matrix Multiplication with C++ AMP

    - by Daniel Moth
    As part of our API tour of C++ AMP, we looked recently at parallel_for_each. I ended that post by saying we would revisit parallel_for_each after introducing array and array_view. Now is the time, so this is part 2 of parallel_for_each, and also a post that brings together everything we've seen until now. The code for serial and accelerated Consider a naïve (or brute force) serial implementation of matrix multiplication  0: void MatrixMultiplySerial(std::vector<float>& vC, const std::vector<float>& vA, const std::vector<float>& vB, int M, int N, int W) 1: { 2: for (int row = 0; row < M; row++) 3: { 4: for (int col = 0; col < N; col++) 5: { 6: float sum = 0.0f; 7: for(int i = 0; i < W; i++) 8: sum += vA[row * W + i] * vB[i * N + col]; 9: vC[row * N + col] = sum; 10: } 11: } 12: } We notice that each loop iteration is independent from each other and so can be parallelized. If in addition we have really large amounts of data, then this is a good candidate to offload to an accelerator. First, I'll just show you an example of what that code may look like with C++ AMP, and then we'll analyze it. It is assumed that you included at the top of your file #include <amp.h> 13: void MatrixMultiplySimple(std::vector<float>& vC, const std::vector<float>& vA, const std::vector<float>& vB, int M, int N, int W) 14: { 15: concurrency::array_view<const float,2> a(M, W, vA); 16: concurrency::array_view<const float,2> b(W, N, vB); 17: concurrency::array_view<concurrency::writeonly<float>,2> c(M, N, vC); 18: concurrency::parallel_for_each(c.grid, 19: [=](concurrency::index<2> idx) restrict(direct3d) { 20: int row = idx[0]; int col = idx[1]; 21: float sum = 0.0f; 22: for(int i = 0; i < W; i++) 23: sum += a(row, i) * b(i, col); 24: c[idx] = sum; 25: }); 26: } First a visual comparison, just for fun: The beginning and end is the same, i.e. lines 0,1,12 are identical to lines 13,14,26. The double nested loop (lines 2,3,4,5 and 10,11) has been transformed into a parallel_for_each call (18,19,20 and 25). The core algorithm (lines 6,7,8,9) is essentially the same (lines 21,22,23,24). We have extra lines in the C++ AMP version (15,16,17). Now let's dig in deeper. Using array_view and extent When we decided to convert this function to run on an accelerator, we knew we couldn't use the std::vector objects in the restrict(direct3d) function. So we had a choice of copying the data to the the concurrency::array<T,N> object, or wrapping the vector container (and hence its data) with a concurrency::array_view<T,N> object from amp.h – here we used the latter (lines 15,16,17). Now we can access the same data through the array_view objects (a and b) instead of the vector objects (vA and vB), and the added benefit is that we can capture the array_view objects in the lambda (lines 19-25) that we pass to the parallel_for_each call (line 18) and the data will get copied on demand for us to the accelerator. Note that line 15 (and ditto for 16 and 17) could have been written as two lines instead of one: extent<2> e(M, W); array_view<const float, 2> a(e, vA); In other words, we could have explicitly created the extent object instead of letting the array_view create it for us under the covers through the constructor overload we chose. The benefit of the extent object in this instance is that we can express that the data is indeed two dimensional, i.e a matrix. When we were using a vector object we could not do that, and instead we had to track via additional unrelated variables the dimensions of the matrix (i.e. with the integers M and W) – aren't you loving C++ AMP already? Note that the const before the float when creating a and b, will result in the underling data only being copied to the accelerator and not be copied back – a nice optimization. A similar thing is happening on line 17 when creating array_view c, where we have indicated that we do not need to copy the data to the accelerator, only copy it back. The kernel dispatch On line 18 we make the call to the C++ AMP entry point (parallel_for_each) to invoke our parallel loop or, as some may say, dispatch our kernel. The first argument we need to pass describes how many threads we want for this computation. For this algorithm we decided that we want exactly the same number of threads as the number of elements in the output matrix, i.e. in array_view c which will eventually update the vector vC. So each thread will compute exactly one result. Since the elements in c are organized in a 2-dimensional manner we can organize our threads in a two-dimensional manner too. We don't have to think too much about how to create the first argument (a grid) since the array_view object helpfully exposes that as a property. Note that instead of c.grid we could have written grid<2>(c.extent) or grid<2>(extent<2>(M, N)) – the result is the same in that we have specified M*N threads to execute our lambda. The second argument is a restrict(direct3d) lambda that accepts an index object. Since we elected to use a two-dimensional extent as the first argument of parallel_for_each, the index will also be two-dimensional and as covered in the previous posts it represents the thread ID, which in our case maps perfectly to the index of each element in the resulting array_view. The kernel itself The lambda body (lines 20-24), or as some may say, the kernel, is the code that will actually execute on the accelerator. It will be called by M*N threads and we can use those threads to index into the two input array_views (a,b) and write results into the output array_view ( c ). The four lines (21-24) are essentially identical to the four lines of the serial algorithm (6-9). The only difference is how we index into a,b,c versus how we index into vA,vB,vC. The code we wrote with C++ AMP is much nicer in its indexing, because the dimensionality is a first class concept, so you don't have to do funny arithmetic calculating the index of where the next row starts, which you have to do when working with vectors directly (since they store all the data in a flat manner). I skipped over describing line 20. Note that we didn't really need to read the two components of the index into temporary local variables. This mostly reflects my personal choice, in some algorithms to break down the index into local variables with names that make sense for the algorithm, i.e. in this case row and col. In other cases it may i,j,k or x,y,z, or M,N or whatever. Also note that we could have written line 24 as: c(idx[0], idx[1])=sum  or  c(row, col)=sum instead of the simpler c[idx]=sum Targeting a specific accelerator Imagine that we had more than one hardware accelerator on a system and we wanted to pick a specific one to execute this parallel loop on. So there would be some code like this anywhere before line 18: vector<accelerator> accs = MyFunctionThatChoosesSuitableAccelerators(); accelerator acc = accs[0]; …and then we would modify line 18 so we would be calling another overload of parallel_for_each that accepts an accelerator_view as the first argument, so it would become: concurrency::parallel_for_each(acc.default_view, c.grid, ...and the rest of your code remains the same… how simple is that? Comments about this post by Daniel Moth welcome at the original blog.

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  • DBA Best Practices - A Blog Series: Episode 1 - Backups

    - by Argenis
      This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups had completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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  • MySQL port 3306 blocked in csf yet can still telnet to port 3306 from external host

    - by Neek
    We have a Centos 6 VPS that was recently migrated to a new machine within the same web hosting company. It's running WHM/cPanel and has csf/lfd installed. csf is set up with mostly vanilla config. I'm no iptables expert, csf has not let me down before. If a port isn't in the TCP_IN list, it should be blocked on the firewall by iptables. My problem is that I can telnet to port 3306 from an external host, yet I think iptables ought to be blocking 3306 because of csf's rules. We are now failing a security check because of this open port. (this output is obfuscated to protect the innocent: www.ourhost.com is the host with the firewall problem) [root@nickfenwick log]# telnet www.ourhost.com 3306 Trying 158.255.45.107... Connected to www.ourhost.com. Escape character is '^]'. HHost 'nickfenwick.com' is not allowed to connect to this MySQL serverConnection closed by foreign host. So the connection is established, and MySQL refuses the connection due to its configuration. I need the network connection to be refused at the firewall level, before it reaches MySQL. Using WHM's csf web UI I can see 'Firewall Configuration' includes a fairly sensible TCP_IN line: TCP_IN: 20,21,22,25,53,80,110,143,222,443,465,587,993,995,2077,2078,2082,2083,2086,2087,2095,2096,8080 (lets ignore that I could trim that a little for now, my concern is that 3306 is not listed in that list) When csf is restarted it logs the usual slew of output as it sets up iptables rules, for example what looks like it blocking all traffic and then allowing specific ports like SSH on 22: [cut] DROP all opt -- in * out * 0.0.0.0/0 -> 0.0.0.0/0 [cut] ACCEPT tcp opt -- in !lo out * 0.0.0.0/0 -> 0.0.0.0/0 state NEW tcp dpt:22 [cut] I can see that iptables is running, service iptables status returns a long list of firewall rules. Here is my Chain INPUT section from service iptables status, hopefully that's enough to show how the firewall is configured. Table: filter Chain INPUT (policy DROP) num target prot opt source destination 1 acctboth all -- 0.0.0.0/0 0.0.0.0/0 2 ACCEPT tcp -- 217.112.88.10 0.0.0.0/0 tcp dpt:53 3 ACCEPT udp -- 217.112.88.10 0.0.0.0/0 udp dpt:53 4 ACCEPT tcp -- 217.112.88.10 0.0.0.0/0 tcp spt:53 5 ACCEPT udp -- 217.112.88.10 0.0.0.0/0 udp spt:53 6 ACCEPT tcp -- 8.8.4.4 0.0.0.0/0 tcp dpt:53 7 ACCEPT udp -- 8.8.4.4 0.0.0.0/0 udp dpt:53 8 ACCEPT tcp -- 8.8.4.4 0.0.0.0/0 tcp spt:53 9 ACCEPT udp -- 8.8.4.4 0.0.0.0/0 udp spt:53 10 ACCEPT tcp -- 8.8.8.8 0.0.0.0/0 tcp dpt:53 11 ACCEPT udp -- 8.8.8.8 0.0.0.0/0 udp dpt:53 12 ACCEPT tcp -- 8.8.8.8 0.0.0.0/0 tcp spt:53 13 ACCEPT udp -- 8.8.8.8 0.0.0.0/0 udp spt:53 14 LOCALINPUT all -- 0.0.0.0/0 0.0.0.0/0 15 ACCEPT all -- 0.0.0.0/0 0.0.0.0/0 16 INVALID tcp -- 0.0.0.0/0 0.0.0.0/0 17 ACCEPT all -- 0.0.0.0/0 0.0.0.0/0 state RELATED,ESTABLISHED 18 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:20 19 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:21 20 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:22 21 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:25 22 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:53 23 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:80 24 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:110 25 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:143 26 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:222 27 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:443 28 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:465 29 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:587 30 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:993 31 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:995 32 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2077 33 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2078 34 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2082 35 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2083 36 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2086 37 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2087 38 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2095 39 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:2096 40 ACCEPT tcp -- 0.0.0.0/0 0.0.0.0/0 state NEW tcp dpt:8080 41 ACCEPT udp -- 0.0.0.0/0 0.0.0.0/0 state NEW udp dpt:20 42 ACCEPT udp -- 0.0.0.0/0 0.0.0.0/0 state NEW udp dpt:21 43 ACCEPT udp -- 0.0.0.0/0 0.0.0.0/0 state NEW udp dpt:53 44 ACCEPT udp -- 0.0.0.0/0 0.0.0.0/0 state NEW udp dpt:222 45 ACCEPT udp -- 0.0.0.0/0 0.0.0.0/0 state NEW udp dpt:8080 46 ACCEPT icmp -- 0.0.0.0/0 0.0.0.0/0 icmp type 8 47 ACCEPT icmp -- 0.0.0.0/0 0.0.0.0/0 icmp type 0 48 ACCEPT icmp -- 0.0.0.0/0 0.0.0.0/0 icmp type 11 49 ACCEPT icmp -- 0.0.0.0/0 0.0.0.0/0 icmp type 3 50 LOGDROPIN all -- 0.0.0.0/0 0.0.0.0/0 What's the next thing to check?

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  • &lt;%: %&gt;, HtmlEncode, IHtmlString and MvcHtmlString

    - by Shaun
    One of my colleague and friend, Robin is playing and struggling with the ASP.NET MVC 2 on a project these days while I’m struggling with a annoying client. Since it’s his first time to use ASP.NET MVC he was meetings with a lot of problem and I was very happy to share my experience to him. Yesterday he asked me when he attempted to insert a <br /> element into his page he found that the page was rendered like this which is bad. He found his <br /> was shown as a part of the string rather than creating a new line. After checked a bit in his code I found that it’s because he utilized a new ASP.NET markup supported in .NET 4.0 – “<%: %>”. If you have been using ASP.NET MVC 1 or in .NET 3.5 world it would be very common that using <%= %> to show something on the page from the backend code. But when you do it you must ensure that the string that are going to be displayed should be Html-safe, which means all the Html markups must be encoded. Otherwise this might cause an XSS (cross-site scripting) problem. So that you’d better use the code like this below to display anything on the page. In .NET 4.0 Microsoft introduced a new markup to solve this problem which is <%: %>. It will encode the content automatically so that you will no need to check and verify your code manually for the XSS issue mentioned below. But this also means that it will encode all things, include the Html element you want to be rendered. So I changed his code like this and it worked well. After helped him solved this problem and finished a spreadsheet for my boring project I considered a bit more on the <%: %>. Since it will encode all thing why it renders correctly when we use “<%: Html.TextBox(“name”) %>” to show a text box? As you know the Html.TextBox will render a “<input name="name" id="name" type="text"/>” element on the page. If <%: %> will encode everything it should not display a text box. So I dig into the source code of the MVC and found some comments in the class MvcHtmlString. 1: // In ASP.NET 4, a new syntax <%: %> is being introduced in WebForms pages, where <%: expression %> is equivalent to 2: // <%= HttpUtility.HtmlEncode(expression) %>. The intent of this is to reduce common causes of XSS vulnerabilities 3: // in WebForms pages (WebForms views in the case of MVC). This involves the addition of an interface 4: // System.Web.IHtmlString and a static method overload System.Web.HttpUtility::HtmlEncode(object). The interface 5: // definition is roughly: 6: // public interface IHtmlString { 7: // string ToHtmlString(); 8: // } 9: // And the HtmlEncode(object) logic is roughly: 10: // - If the input argument is an IHtmlString, return argument.ToHtmlString(), 11: // - Otherwise, return HtmlEncode(Convert.ToString(argument)). 12: // 13: // Unfortunately this has the effect that calling <%: Html.SomeHelper() %> in an MVC application running on .NET 4 14: // will end up encoding output that is already HTML-safe. As a result, we're changing out HTML helpers to return 15: // MvcHtmlString where appropriate. <%= Html.SomeHelper() %> will continue to work in both .NET 3.5 and .NET 4, but 16: // changing the return types to MvcHtmlString has the added benefit that <%: Html.SomeHelper() %> will also work 17: // properly in .NET 4 rather than resulting in a double-encoded output. MVC developers in .NET 4 will then be able 18: // to use the <%: %> syntax almost everywhere instead of having to remember where to use <%= %> and where to use 19: // <%: %>. This should help developers craft more secure web applications by default. 20: // 21: // To create an MvcHtmlString, use the static Create() method instead of calling the protected constructor. The comment said the encoding rule of the <%: %> would be: If the type of the content is IHtmlString it will NOT encode since the IHtmlString indicates that it’s Html-safe. Otherwise it will use HtmlEncode to encode the content. If we check the return type of the Html.TextBox method we will find that it’s MvcHtmlString, which was implemented the IHtmlString interface dynamically. That is the reason why the “<input name="name" id="name" type="text"/>” was not encoded by <%: %>. So if we want to tell ASP.NET MVC, or I should say the ASP.NET runtime that the content is Html-safe and no need, or should not be encoded we can convert the content into IHtmlString. So another resolution would be like this. Also we can create an extension method as well for better developing experience. 1: using System; 2: using System.Collections.Generic; 3: using System.Linq; 4: using System.Web; 5: using System.Web.Mvc; 6:  7: namespace ShaunXu.Blogs.IHtmlStringIssue 8: { 9: public static class Helpers 10: { 11: public static MvcHtmlString IsHtmlSafe(this string content) 12: { 13: return MvcHtmlString.Create(content); 14: } 15: } 16: } Then the view would be like this. And the page rendered correctly.         Summary In this post I explained a bit about the new markup in .NET 4.0 – <%: %> and its usage. I also explained a bit about how to control the page content, whether it should be encoded or not. We can see the ASP.NET MVC gives us more points to control the web pages.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

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  • Modifying a gedit syntax highlighting file

    - by Oscar Saleta Reig
    I am trying to change a highlighting file from Gedit. I have modified the file /usr/share/gtksourceview-3.0/language-specs/fortran.lang because I want to change the cases in which the editor takes a statement as a comment. The problem I have is that when I choose the new highlighting scheme nothing highlights, it just remains as plain text. The file fortran.lang was opened with su permissions and I just copy-pasted everything into a new Gedit file and later saved it as fortran_enhanced.lang in the same folder. The changes I've done to the original file are these: Original fortran.lang file: <language id="fortran" _name="Fortran 95" version="2.0" _section="Sources"> <metadata> <property name="mimetypes">text/x-fortran</property> <property name="globs">*.f;*.f90;*.f95;*.for</property> <property name="line-comment-start">!</property> </metadata> <styles> <style id="comment" _name="Comment" map-to="def:comment"/> <style id="floating-point" _name="Floating Point" map-to="def:floating-point"/> <style id="keyword" _name="Keyword" map-to="def:keyword"/> <style id="intrinsic" _name="Intrinsic function" map-to="def:builtin"/> <style id="boz-literal" _name="BOZ Literal" map-to="def:base-n-integer"/> <style id="decimal" _name="Decimal" map-to="def:decimal"/> <style id="type" _name="Data Type" map-to="def:type"/> </styles> <default-regex-options case-sensitive="false"/> <definitions> <!-- Note: contains an hack to avoid considering ^COMMON a comment --> <context id="line-comment" style-ref="comment" end-at-line-end="true" class="comment" class-disabled="no-spell-check"> <start>!|(^[Cc](\b|[^OoAaYy]))</start> <include> <context ref="def:escape"/> <context ref="def:in-line-comment"/> </include> </context> (...) Modified fortran_enhanced.lang file: <!-- Note: changed language id and name --> <language id="fortran_enhanced" _name="Fortran 95 2.0" version="2.0" _section="Sources"> <metadata> <property name="mimetypes">text/x-fortran</property> <!-- Note: removed *.f and *.for from file extensions --> <property name="globs">*.f90;*.f95;</property> <property name="line-comment-start">!</property> </metadata> <styles> <style id="comment" _name="Comment" map-to="def:comment"/> <style id="floating-point" _name="Floating Point" map-to="def:floating-point"/> <style id="keyword" _name="Keyword" map-to="def:keyword"/> <style id="intrinsic" _name="Intrinsic function" map-to="def:builtin"/> <style id="boz-literal" _name="BOZ Literal" map-to="def:base-n-integer"/> <style id="decimal" _name="Decimal" map-to="def:decimal"/> <style id="type" _name="Data Type" map-to="def:type"/> </styles> <default-regex-options case-sensitive="false"/> <definitions> <!-- Note: I want comments only beginning with !, not C --> <context id="line-comment" style-ref="comment" end-at-line-end="true" class="comment" class-disabled="no-spell-check"> <start>!</start> <include> <context ref="def:escape"/> <context ref="def:in-line-comment"/> </include> </context> (...) I have read this question [ Custom gedit Syntax Highlighting for Dummies? ] and I tried to make the new fortran_enhanced.lang file readable with $ cd /usr/share/gtksourceview-3.0/language-specs $ sudo chmod 0644 fortran_enhanced.lang but it doesn't seem that made some difference. I have to say that I have never done a thing like this before and I don't even understand most of the language file, so I am open to every criticism, as I have been guided purely by intuition. Thank you in advanced!

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  • Software: Launching League of Legends spectator mode from Command Line (Mac)

    - by Alex Popov
    Background: tl;dr at the end League of Legends has a spectator mode, in which you can watch someone else's game (essentially a replay) with a 3 minute delay. Popular LoL website OP.GG has figured out a clever way of hosting these spectator games on their own servers, thereby making them replayable, as opposed to only being available while the game is on (as Riot does it). If you request a replay from OP.GG, it sends a batch file which looks for where the League is situated and then the magic happens: @start "" "League of Legends.exe" "8394" "LoLLauncher.exe" "" "spectator fspectate.op.gg:4081 tjJbtRLQ/HMV7HuAxWV0XsXoRB4OmFBr 1391881421 NA1" This works fine on Windows. I'm trying to get it to work on Mac (which has an official client). First I tried running the same command by hand, (split for convenience) /Applications/ ... /LeagueOfLegends.app/ ... /LeagueofLegends 8393 LoLLauncher \ /Applications/ ... /LolClient spectator fspectate.op.gg:4081 tjJbtRLQ/HMV7HuAxWV0XsXoRB4OmFBr 1391881421 NA1 Running this, however, just starts the LoLLauncher, which closes all the active League processes. The exactly same thing happens if I just call /Applications/ ... /LeagueOfLegends.app/ ... /LeagueofLegends Next I tried seeing what actually happens when Spectator mode is initiated so I ran $ ps -axf | grep -i lol which showed UID PID PPID C STIME TTY TIME CMD 503 3085 1 0 Wed02pm ?? 0:00.00 (LolClient) 503 24607 1 0 9:19am ?? 0:00.98 /Applications/League of Legends.app/Contents/LOL/RADS/system/UserKernel.app/Contents/MacOS/UserKernel updateandrun lol_launcher LoLLauncher.app 503 24610 24607 0 9:19am ?? 1:08.76 /Applications/League of Legends.app/Contents/LoL/RADS/projects/lol_launcher/releases/0.0.0.122/deploy/LoLLauncher.app/Contents/MacOS/LoLLauncher 503 24611 24610 0 9:19am ?? 1:23.02 /Applications/League of Legends.app/Contents/LoL/RADS/projects/lol_air_client/releases/0.0.0.127/deploy/bin/LolClient -runtime .\ -nodebug META-INF\AIR\application.xml .\ -- 8393 503 24927 24610 0 9:44am ?? 0:03.37 /Applications/League of Legends.app/Contents/LoL/RADS/solutions/lol_game_client_sln/releases/0.0.0.117/deploy/LeagueOfLegends.app/Contents/MacOS/LeagueofLegends 8394 LoLLauncher /Applications/League of Legends.app/Contents/LoL/RADS/projects/lol_air_client/releases/0.0.0.127/deploy/bin/LolClient spectator 216.133.234.17:8088 Yn1oMX/n3LpXNebibzUa1i3Z+s2HV0ul 1400781241 NA1 Of Interest: there is (LolClient) which I cannot kill by it's PID. UserKernel updateandrun lol_launcher LoLLauncher.app is launched first. LoLLauncher is launched by the UserKernel (as we can see from the PPID) The very long command (PID: 24927) is how Spectator mode is launched, and is also launched by UserKernel. Spectator mode is launched in exactly the same way that the OP.GG .bat wanted to, with the only difference that Spectator mode connects to Riot instead of OP.GG's spectate server. I tried attaching GDB to the LolClient, but I couldn't get anything meaningful from it since it's an Adobe AIR application (and I've never used GDB with code other than mine own). Next I ran dtruss -a -b 100m -f -p $PID on everything I could think of: the LolClient, the LolLauncher and the UserKernel and skimmed the half a million lines produced. I found stuff like the GET request used to get the information of the game to spectate, but I could not see any launch of the equivalent of League of Legends.exe with spectator options. Finally, I ran lsof | grep -i lol to see if anything else was opened in the process, but didn't find anything that seemed appropriate. Open were UserKernel, LolLauncher, LolClient, Adobe AIR, LeagueofLegends and then Bugsplat, all of which are expected. None of this seemed especially relevant to figuring out how LeagueofLegends was opened into spectator mode. It obviously can be done, since Spectator mode is accessible from within the client. It seems likely that it can be done from the CLI, since Windows can do it and the clients are supposed to equals. Unless I'm missing something in the difference between how UNIX and Windows handle CLI application launches. My question is if there are any other things I can try to figure out how to launch Spectator mode myself. tl;dr: Trying to get into spectator mode from the CLI. It's possible on Windows (see first code block) but it just restarts League on Mac. What else can I try to find what call is made, and how to reproduce it? PS: Please let me know how I can improve this question or its formatting, I'd love to use StackOverflow/SuperUser, but as the guys said on the podcast this week (Ep. 59) it's very intimidating. Sorry for posting this on StackOverflow the first time :(

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  • Agilist, Heal Thyself!

    - by Dylan Smith
    I’ve been meaning to blog about a great experience I had earlier in the year at Prairie Dev Con Calgary.  Myself and Steve Rogalsky did a session that we called “Agilist, Heal Thyself!”.  We used a format that was new to me, but that Steve had seen used at another conference.  What we did was start by asking the audience to give us a list of challenges they had had when adopting agile.  We wrote them all down, then had everybody vote on the most interesting ones.  Then we split into two groups, and each group was assigned one of the agile challenges.  We had 20 minutes to discuss the challenge, and suggest solutions or approaches to improve things.  At the end of the 20 minutes, each of the groups gave a brief summary of their discussion and learning's, then we mixed up the groups and repeated with another 2 challenges. The 2 groups I was part of had some really interesting discussions, and suggestions: Unfinished Stories at the end of Sprints The first agile challenge we tackled, was something that every single Scrum team I have worked with has struggled with.  What happens when you get to the end of a Sprint, and there are some stories that are only partially completed.  The team in question was getting very de-moralized as they felt that every Sprint was a failure as they never had a set of fully completed stories. How do you avoid this? and/or what do you do when it happens? There were 2 pieces of advice that were well received: 1. Try to bring stories to completion before starting new ones.  This is advice I give all my Scrum teams.  If you have a 3-week sprint, what happens all too often is you get to the end of week 2, and a lot of stories are almost done; but almost none are completely done.  This is a Bad Thing.  I encourage the teams I work with to only start a new story as a very last resort.  If you finish your task look at the stories in progress and see if there’s anything you can do to help before moving onto a new story.  In the daily standup, put a focus on seeing what stories got completed yesterday, if a few days go by with none getting completed, be sure this fact is visible to the team and do something about it.  Something I’ve been doing recently is introducing WIP (Work In Progress) limits while using Scrum.  My current team has 2-week sprints, and we usually have about a dozen or stories in a sprint.  We instituted a WIP limit of 4 stories.  If 4 stories have been started but not finished then nobody is allowed to start new stories.  This made it obvious very quickly that our QA tasks were our bottleneck (we have 4 devs, but only 1.5 testers).  The WIP limit forced the developers to start to pickup QA tasks before moving onto the next dev tasks, and we ended our sprints with many more stories completely finished than we did before introducing WIP limits. 2. Rather than using time-boxed sprints, why not just do away with them altogether and go to a continuous flow type approach like KanBan.  Limit WIP to keep things under control, but don’t have a fixed time box at the end of which all tasks are supposed to be done.  This eliminates the problem almost entirely.  At some points in the project (releases) you need to be able to burn down all the half finished stories to get a stable release build, but this probably occurs less often than every sprint, and there are alternative approaches to achieve it using branching strategies rather than forcing your team to try to get to Zero WIP every 2-weeks (e.g. when you are ready for a release, create a new branch for any new stories, but finish all existing stories in the current branch and release it). Trying to Introduce Agile into a team with previous Bad Agile Experiences One of the agile adoption challenges somebody described, was he was in a leadership role on a team he had recently joined – lets call him Dave.  This team was currently very waterfall in their ALM process, but they were about to start on a new green-field project.  Dave wanted to use this new project as an opportunity to do things the “right way”, using an Agile methodology like Scrum, adopting TDD, automated builds, proper branching strategies, etc.  The problem he was facing is everybody else on the team had previously gone through an “Agile Adoption” that was a horrible failure.  Dave blamed this failure on the consultant brought in previously to lead this agile transition, but regardless of the reason, the team had very negative feelings towards agile, and was very resistant to trying it out again.  Dave possibly had the authority to try to force the team to adopt Agile practices, but we all know that doesn’t work very well.  What was Dave to do? Ultimately, the best advice was to question *why* did Dave want to adopt all these various practices. Rather than trying to convince his team that these were the “right way” to run a dev project, and trying to do a Big Bang approach to introducing change.  He would be better served by identifying problems the team currently faces, have a discussion with the team to get everybody to agree that specific problems existed, then have an open discussion about ways to address those problems.  This way Dave could incrementally introduce agile practices, and he doesn’t even need to identify them as “agile” practices if he doesn’t want to.  For example, when we discussed with Dave, he said probably the teams biggest problem was long periods without feedback from users, then finding out too late that the software is not going to meet their needs.  Rather than Dave jumping right to introducing Scrum and all it entails, it would be easier to get buy-in from team if he framed it as a discussion of existing problems, and brainstorming possible solutions.  And possibly most importantly, don’t try to do massive changes all at once with a team that has not bought-into those changes.  Taking an incremental approach has a greater chance of success. I see something similar in my day job all the time too.  Clients who for one reason or another claim to not be fans of agile (or not ready for agile yet).  But then they go on to ask me to help them get shorter feedback cycles, quicker delivery cycles, iterative development processes, etc.  It’s kind of funny at times, sometimes you just need to phrase the suggestions in terms they are using and avoid the word “agile”. PS – I haven’t blogged all that much over the past couple of years, but in an attempt to motivate myself, a few of us have accepted a blogger challenge.  There’s 6 of us who have all put some money into a pool, and the agreement is that we each need to blog at least once every 2-weeks.  The first 2-week period that we miss we’re eliminated.  Last person standing gets the money.  So expect at least one blog post every couple of weeks for the near future (I hope!).  And check out the blogs of the other 5 people in this blogger challenge: Steve Rogalsky: http://winnipegagilist.blogspot.ca Aaron Kowall: http://www.geekswithblogs.net/caffeinatedgeek Tyler Doerkson: http://blog.tylerdoerksen.com David Alpert: http://www.spinthemoose.com Dave White: http://www.agileramblings.com (note: site not available yet.  should be shortly or he owes me some money!)

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  • Code contracts and inheritance

    - by DigiMortal
    In my last posting about code contracts I introduced you how to force code contracts to classes through interfaces. In this posting I will go step further and I will show you how code contracts work in the case of inherited classes. As a first thing let’s take a look at my interface and code contracts. [ContractClass(typeof(ProductContracts))] public interface IProduct {     int Id { get; set; }     string Name { get; set; }     decimal Weight { get; set; }     decimal Price { get; set; } }   [ContractClassFor(typeof(IProduct))] internal sealed class ProductContracts : IProduct {     private ProductContracts() { }       int IProduct.Id     {         get         {             return default(int);         }         set         {             Contract.Requires(value > 0);         }     }       string IProduct.Name     {         get         {             return default(string);         }         set         {             Contract.Requires(!string.IsNullOrWhiteSpace(value));             Contract.Requires(value.Length <= 25);         }     }       decimal IProduct.Weight     {         get         {             return default(decimal);         }         set         {             Contract.Requires(value > 3);             Contract.Requires(value < 100);         }     }       decimal IProduct.Price     {         get         {             return default(decimal);         }         set         {             Contract.Requires(value > 0);             Contract.Requires(value < 100);         }     } } And here is the product class that inherits IProduct interface. public class Product : IProduct {     public int Id { get; set; }     public string Name { get; set; }     public virtual decimal Weight { get; set; }     public decimal Price { get; set; } } if we run this code and violate the code contract set to Id we will get ContractException. public class Program {     static void Main(string[] args)     {         var product = new Product();         product.Id = -100;     } }   Now let’s make Product to be abstract class and let’s define new class called Food that adds one more contract to Weight property. public class Food : Product {     public override decimal Weight     {         get         {             return base.Weight;         }         set         {             Contract.Requires(value > 1);             Contract.Requires(value < 10);               base.Weight = value;         }     } } Now we should have the following rules at place for Food: weight must be greater than 1, weight must be greater than 3, weight must be less than 100, weight must be less than 10. Interesting part is what happens when we try to violate the lower and upper limits of Food weight. To see what happens let’s try to violate rules #2 and #4. Just comment one of the last lines out in the following method to test another assignment. public class Program {     static void Main(string[] args)     {         var food = new Food();         food.Weight = 12;         food.Weight = 2;     } } And here are the results as pictures to see where exceptions are thrown. Click on images to see them at original size. Violation of lower limit. Violation of upper limit. As you can see for both violations we get ContractException like expected. Code contracts inheritance is powerful and at same time dangerous feature. Although you can always narrow down the conditions that come from more general classes it is possible to define impossible or conflicting contracts at different points in inheritance hierarchy.

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  • SQL Server Developer Tools &ndash; Codename Juneau vs. Red-Gate SQL Source Control

    - by Ajarn Mark Caldwell
    So how do the new SQL Server Developer Tools (previously code-named Juneau) stack up against SQL Source Control?  Read on to find out. At the PASS Community Summit a couple of weeks ago, it was announced that the previously code-named Juneau software would be released under the name of SQL Server Developer Tools with the release of SQL Server 2012.  This replacement for Database Projects in Visual Studio (also known in a former life as Data Dude) has some great new features.  I won’t attempt to describe them all here, but I will applaud Microsoft for making major improvements.  One of my favorite changes is the way database elements are broken down.  Previously every little thing was in its own file.  For example, indexes were each in their own file.  I always hated that.  Now, SSDT uses a pattern similar to Red-Gate’s and puts the indexes and keys into the same file as the overall table definition. Of course there are really cool features to keep your database model in sync with the actual source scripts, and the rename refactoring feature is now touted as being more than just a search and replace, but rather a “semantic-aware” search and replace.  Funny, it reminds me of SQL Prompt’s Smart Rename feature.  But I’m not writing this just to criticize Microsoft and argue that they are late to the party with this feature set.  Instead, I do see it as a viable alternative for folks who want all of their source code to be version controlled, but there are a couple of key trade-offs that you need to know about when you choose which tool set to use. First, the basics Both tool sets integrate with a wide variety of source control systems including the most popular: Subversion, GIT, Vault, and Team Foundation Server.  Both tools have integrated functionality to produce objects to upgrade your target database when you are ready (DACPACs in SSDT, integration with SQL Compare for SQL Source Control).  If you regularly live in Visual Studio or the Business Intelligence Development Studio (BIDS) then SSDT will likely be comfortable for you.  Like BIDS, SSDT is a Visual Studio Project Type that comes with SQL Server, and if you don’t already have Visual Studio installed, it will install the shell for you.  If you already have Visual Studio 2010 installed, then it will just add this as an available project type.  On the other hand, if you regularly live in SQL Server Management Studio (SSMS) then you will really enjoy the SQL Source Control integration from within SSMS.  Both tool sets store their database model in script files.  In SSDT, these are on your file system like other source files; in SQL Source Control, these are stored in the folder structure in your source control system, and you can always GET them to your file system if you want to browse them directly. For me, the key differentiating factors are 1) a single, unified check-in, and 2) migration scripts.  How you value those two features will likely make your decision for you. Unified Check-In If you do a continuous-integration (CI) style of development that triggers an automated build with unit testing on every check-in of source code, and you use Visual Studio for the rest of your development, then you will want to really consider SSDT.  Because it is just another project in Visual Studio, it can be added to your existing Solution, and you can then do a complete, or unified single check-in of all changes whether they are application or database changes.  This is simply not possible with SQL Source Control because it is in a different development tool (SSMS instead of Visual Studio) and there is no way to do one unified check-in between the two.  You CAN do really fast back-to-back check-ins, but there is the possibility that the automated build that is triggered from the first check-in will cause your unit tests to fail and the CI tool to report that you broke the build.  Of course, the automated build that is triggered from the second check-in which contains the “other half” of your changes should pass and so the amount of time that the build was broken may be very, very short, but if that is very, very important to you, then SQL Source Control just won’t work; you’ll have to use SSDT. Refactoring and Migrations If you work on a mature system, or on a not-so-mature but also not-so-well-designed system, where you want to refactor the database schema as you go along, but you can’t have data suddenly disappearing from your target system, then you’ll probably want to go with SQL Source Control.  As I wrote previously, there are a number of changes which you can make to your database that the comparison tools (both from Microsoft and Red Gate) simply cannot handle without the possibility (or probability) of data loss.  Currently, SSDT only offers you the ability to inject PRE and POST custom deployment scripts.  There is no way to insert your own script in the middle to override the default behavior of the tool.  In version 3.0 of SQL Source Control (Early Access version now available) you have that ability to create your own custom migration script to take the place of the commands that the tool would have done, and ensure the preservation of your data.  Or, even if the default tool behavior would have worked, but you simply know a better way then you can take control and do things your way instead of theirs. You Decide In the environment I work in, our automated builds are not triggered off of check-ins, but off of the clock (currently once per night) and so there is no point at which the automated build and unit tests will be triggered without having both sides of the development effort already checked-in.  Therefore having a unified check-in, while handy, is not critical for us.  As for migration scripts, these are critically important to us.  We do a lot of new development on systems that have already been in production for years, and it is not uncommon for us to need to do a refactoring of the database.  Because of the maturity of the existing system, that often involves data migrations or other additional SQL tasks that the comparison tools just can’t detect on their own.  Therefore, the ability to create a custom migration script to override the tool’s default behavior is very important to us.  And so, you can see why we will continue to use Red Gate SQL Source Control for the foreseeable future.

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  • Who could ask for more with LESS CSS? (Part 3 of 3&ndash;Clrizr)

    - by ToString(theory);
    Welcome back!  In the first two posts in this series, I covered some of the awesome features in CSS precompilers such as SASS and LESS, as well as how to get an initial project setup up and running in ASP.Net MVC 4. In this post, I will cover an actual advanced example of using LESS in a project, and show some of the great productivity features we gain from its usage. Introduction In the first post, I mentioned two subjects that I will be using in this example – constants, and color functions.  I’ve always enjoyed using online color scheme utilities such as Adobe Kuler or Color Scheme Designer to come up with a scheme based off of one primary color.  Using these tools, and requesting a complementary scheme you can get a couple of shades of your primary color, and a couple of shades of a complementary/accent color to display. Because there is no way in regular css to do color operations or store variables, there was no way to accomplish something like defining a primary color, and have a site theme cascade off of that.  However with tools such as LESS, that impossibility becomes a reality!  So, if you haven’t guessed it by now, this post is on the creation of a plugin/module/less file to drop into your project, plugin one color, and have your primary theme cascade from it.  I only went through the trouble of creating a module for getting Complementary colors.  However, it wouldn’t be too much trouble to go through other options such as Triad or Monochromatic to get a module that you could use off of that. Step 1 – Analysis I decided to mimic Adobe Kuler’s Complementary theme algorithm as I liked its simplicity and aesthetics.  Color Scheme Designer is great, but I do believe it can give you too many color options, which can lead to chaos and overload.  The first thing I had to check was if the complementary values for the color schemes were actually hues rotated by 180 degrees at all times – they aren’t.  Apparently Adobe applies some variance to the complementary colors to get colors that are actually more aesthetically appealing to users.  So, I opened up Excel and began to plot complementary hues based on rotation in increments of 10: Long story short, I completed the same calculations for Hue, Saturation, and Lightness.  For Hue, I only had to record the Complementary hue values, however for saturation and lightness, I had to record the values for ALL of the shades.  Since the functions were too complicated to put into LESS since they aren’t constant/linear, but rather interval functions, I instead opted to extrapolate the HSL values using the trendline function for each major interval, onto intervals of spacing 1. For example, using the hue extraction, I got the following values: Interval Function 0-60 60-140 140-270 270-360 Saturation and Lightness were much worse, but in the end, I finally had functions for all of the intervals, and then went the route of just grabbing each shades value in intervals of 1.  Step 2 – Mapping I declared variable names for each of these sections as something that shouldn’t ever conflict with a variable someone would define in their own file.  After I had each of the values, I extracted the values and put them into files of their own for hue variables, saturation variables, and lightness variables…  Example: /*HUE CONVERSIONS*/@clrizr-hue-source-0deg: 133.43;@clrizr-hue-source-1deg: 135.601;@clrizr-hue-source-2deg: 137.772;@clrizr-hue-source-3deg: 139.943;@clrizr-hue-source-4deg: 142.114;.../*SATURATION CONVERSIONS*/@clrizr-saturation-s2SV0px: 0;@clrizr-saturation-s2SV1px: 0;@clrizr-saturation-s2SV2px: 0;@clrizr-saturation-s2SV3px: 0;@clrizr-saturation-s2SV4px: 0;.../*LIGHTNESS CONVERSIONS*/@clrizr-lightness-s2LV0px: 30;@clrizr-lightness-s2LV1px: 31;@clrizr-lightness-s2LV2px: 32;@clrizr-lightness-s2LV3px: 33;@clrizr-lightness-s2LV4px: 34;...   In the end, I have 973 lines of mapping/conversion from source HSL to shade HSL for two extra primary shades, and two complementary shades. The last bit of the work was the file to compose each of the shades from these mappings. Step 3 – Clrizr Mapper The final step was the hardest to overcome as I was still trying to understand LESS to its fullest extent.  Imports As mentioned previously, I had separated the HSL mappings into different files, so the first necessary step is to import those for use into the Clrizr plugin: @import url("hue.less");@import url("saturation.less");@import url("lightness.less"); Extract Component Values For Each Shade Next, I extracted the necessary information for each shade HSL before shade composition: @clrizr-input-saturation: 1px+floor(saturation(@clrizr-input))-1;@clrizr-input-lightness: 1px+floor(lightness(@clrizr-input))-1; @clrizr-complementary-hue: formatstring("clrizr-hue-source-{0}", ceil(hue(@clrizr-input))); @clrizr-primary-2-saturation: formatstring("clrizr-saturation-s2SV{0}",@clrizr-input-saturation);@clrizr-primary-1-saturation: formatstring("clrizr-saturation-s1SV{0}",@clrizr-input-saturation);@clrizr-complementary-1-saturation: formatstring("clrizr-saturation-c1SV{0}",@clrizr-input-saturation); @clrizr-primary-2-lightness: formatstring("clrizr-lightness-s2LV{0}",@clrizr-input-lightness);@clrizr-primary-1-lightness: formatstring("clrizr-lightness-s1LV{0}",@clrizr-input-lightness);@clrizr-complementary-1-lightness: formatstring("clrizr-lightness-c1LV{0}",@clrizr-input-lightness); Here, you can see a couple of odd things…  On the first line, I am using operations to add units to the saturation and lightness.  This is due to some limitations in the operations that would give me saturation or lightness in %, which can’t be in a variable name.  So, I use first add 1px to it, which casts the result of the following functions as px instead of %, and then at the end, I remove that pixel.  You can also see here the formatstring method which is exactly what it sounds like – something like String.Format(string str, params object[] obj). Get Primary & Complementary Shades Now that I have components for each of the different shades, I can now compose them into each of their pieces.  For this, I use the @@ operator which will look for a variable with the name specified in a string, and then call that variable: @clrizr-primary-2: hsl(hue(@clrizr-input), @@clrizr-primary-2-saturation, @@clrizr-primary-2-lightness);@clrizr-primary-1: hsl(hue(@clrizr-input), @@clrizr-primary-1-saturation, @@clrizr-primary-1-lightness);@clrizr-primary: @clrizr-input;@clrizr-complementary-1: hsl(@@clrizr-complementary-hue, @@clrizr-complementary-1-saturation, @@clrizr-complementary-1-lightness);@clrizr-complementary-2: hsl(@@clrizr-complementary-hue, saturation(@clrizr-input), lightness(@clrizr-input)); That’s is it, for the most part.  These variables now hold the theme for the one input color – @clrizr-input.  However, I have one last addition… Perceptive Luminance Well, after I got the colors, I decided I wanted to also get the best font color that would go on top of it.  Black or white depending on light or dark color.  Now I couldn’t just go with checking the lightness, as that is half the story.  You see, the human eye doesn’t see ALL colors equally well but rather has more cells for interpreting green light compared to blue or red.  So, using the ratio, we can calculate the perceptive luminance of each of the shades, and get the font color that best matches it! @clrizr-perceptive-luminance-ps2: round(1 - ( (0.299 * red(@clrizr-primary-2) ) + ( 0.587 * green(@clrizr-primary-2) ) + (0.114 * blue(@clrizr-primary-2)))/255)*255;@clrizr-perceptive-luminance-ps1: round(1 - ( (0.299 * red(@clrizr-primary-1) ) + ( 0.587 * green(@clrizr-primary-1) ) + (0.114 * blue(@clrizr-primary-1)))/255)*255;@clrizr-perceptive-luminance-ps: round(1 - ( (0.299 * red(@clrizr-primary) ) + ( 0.587 * green(@clrizr-primary) ) + (0.114 * blue(@clrizr-primary)))/255)*255;@clrizr-perceptive-luminance-pc1: round(1 - ( (0.299 * red(@clrizr-complementary-1)) + ( 0.587 * green(@clrizr-complementary-1)) + (0.114 * blue(@clrizr-complementary-1)))/255)*255;@clrizr-perceptive-luminance-pc2: round(1 - ( (0.299 * red(@clrizr-complementary-2)) + ( 0.587 * green(@clrizr-complementary-2)) + (0.114 * blue(@clrizr-complementary-2)))/255)*255; @clrizr-col-font-on-primary-2: rgb(@clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2, @clrizr-perceptive-luminance-ps2);@clrizr-col-font-on-primary-1: rgb(@clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1, @clrizr-perceptive-luminance-ps1);@clrizr-col-font-on-primary: rgb(@clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps, @clrizr-perceptive-luminance-ps);@clrizr-col-font-on-complementary-1: rgb(@clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1, @clrizr-perceptive-luminance-pc1);@clrizr-col-font-on-complementary-2: rgb(@clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2, @clrizr-perceptive-luminance-pc2); Conclusion That’s it!  I have posted a project on clrizr.codePlex.com for this, and included a testing page for you to test out how it works.  Feel free to use it in your own project, and if you have any questions, comments or suggestions, please feel free to leave them here as a comment, or on the contact page!

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  • I Know What I Did This Summer: Put Down Trex Decking

    - by thatjeffsmith
    If you’re wondering why I would bore everyone with my pictures and frequent status updates/tweets from the past week – it’s so I could document the process of refurbishing my deck, or what some would call a porch. When we go to take a vacation, buy a car, do anything – we also read personal blogs to get the real story. So, if you’re curious about what it takes to tackle this sort of project, read on. Skills/Equipment/Manpower We Possessed I took the old decking out by myself. I’m about 230 lbs, more than 6′ tall, and I’m pretty healthy. This took about 8 hours over two afternoons. Three of us put the deck back together. My wife has two engineering degrees. Her father also has two engineering degrees. Lots of brainpower available here. Also, her dad ran the public works department for a country for more than 20 years – so lots and lots of practical experience on hand. We had a compound mitre saw, a skilsaw, 2-3 crowbars, a framing hammer, 3 cordless drills, a corded drill, lots of sawhorses, a power sander, an angle grinder, a 10×10 Coleman canopy tent, a Ford F-150 pickup truck, outdoor speakers and lots of iTunes playlists, plenty of water and cold beer. Why We Did This Our deck was relatively young – it was built in 2005. However, the pressure treated boards must not have been adequately maintained before we bought the house. I had powerwashed the deck every other year and had it stained a few times. The boards just rotted. We’re going to be in the house for a long time, and we wanted something that would look nice and require little maintenance. More bad deck boards The deck boards were in bad shape Things We Learned The two most important things: The hidden fasteners have to be put in JUST right. Wedge them into the grooved board, then bend down the bit that is screwed down. We didn’t do this on the first board and couldn’t get the second board to fit nearly close enough. Watching the official TREX YouTube video helped immensely, and we should have watched that first. When pre-drilling holes for the boards that need screwed down – DO NOT pre-drill through the underlying framing wood. ONLY pre-drill through the TREX itself. The screw won’t seat in the board properly. Instead of sitting down flush with the board, it will stop at the top of the board and just spin. I had to call the the place that sold me the screws to find this out. So about a third of our screws look like crap. If it doesn’t look or feel right – stop everything and pick up your computer or your phone. It’s not right, and it will be much easier to stop and find out why. We didn’t do this, and now I’m going to see every screw that’s not flush with the boards and get upset. Oh well. The Process How much time did it take? Well I spent about 8 hours taking the deck apart. And then the 3 of use spent 8 hours the first day, 10 hours the second day, 8 hours the third, and another 6 hours on the fourth day. That’s like 104 man-hours. We supposedly saved four or five thousand dollars in labor, but don’t do the math here or you might get a bit upset. The main thing is that we got what we wanted, and there won’t be any surprises later. Now for some pictures… This 6”+ pry bar made the destruction of the old deck much easier Most of the joists, once exposed, were OK. This joist wasn’t sitting on ANYTHING before. We think a lazy gas person cut the board to sneak a gas line in. Awesome… These monster lag bolts had to be accounted for when putting in the additional framing The border pattern Sheri wanted to put in required a lot more framing. These were the first boards to go down – we screwed them in as there was no way to attach clips I sat, kicked in the boards, and then drilled these clips in – but my wife was able to go MUCH faster by using her hands to lock the boards in and drill on her knees. I liked locking the board in with my feet when they needed to be ‘encouraged’ to go straight. The first board took FOREVER to go in, but then when we got rolling, we were able to put in a 20′ board in less than 10 minutes. This was end of construction day #2 – we got much further than we thought we would. Ah, the dreaded last 10% – what to do here? Remember those ‘floating’ stringers? Yeah, we fixed that up a bit, too. My wife used a website (and her brain) to calculate exactly how to cut the stringers to give us the rise/run we needed with the proper clearance and all that jazz. The stairs with stringers and toe kicks – this was worth the effort It started raining on us as I screwed down the steps – this we managed to get our shade tent up on the deck to protect us from the rain too The stairs, finished Finished, mostly Good corner shot The top of the stairs Stairs, looking down Celebratory beer In Summary There are a few things we’re not happy with. I think we can fix them up – but later. I have a few things left to finish, rewire the lighting, get the gas grille put back in, and rehang some screen doors. I was expecting this to be a lot worse than it was. If I didn’t have the help, I would have never done it myself. But I’m glad that I did have that help and did do that project. It’s not often you get to spend that kind of qualify time with family and building cool stuff.

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  • First PC Build (Part 1)

    - by Anthony Trudeau
    Originally posted on: http://geekswithblogs.net/tonyt/archive/2014/08/05/157959.aspxA couple of months ago I made the decision to build myself a new computer. The intended use is gaming and for using the last real version of Photoshop. I was motivated by the poor state of console gaming and a simple desire to do something I haven’t done before – build a PC from the ground up. I’ve been using PCs for more than two decades. I’ve replaced a component hear and there, but for the last 10 years or so I’ve only used laptops. Therefore, this article will be written from the perspective of someone familiar with PCs, but completely new at building. I’m not an expert and this is not a definitive guide for building a PC, but I do hope that it encourages you to try it yourself. Component List Research There was a lot of research necessary, because building a PC is completely new to me, and I haven’t kept up with what’s out there. The first thing you want to do is nail down what your goals are. Your goals are going to be driven by what you want to do with your computer and personal choice. Don’t neglect the second one, because if you’re doing this for fun you want to get what you want. In my case, I focused on three things: performance, longevity, and aesthetics. The performance aspect is important for gaming and Photoshop. This will drive what components you get. For example, heavy gaming use is going to drive your choice of graphics card. Longevity is relevant to me, because I don’t want to be changing things out anytime soon for the next hot game. The consequence of performance and longevity is cost. Finally, aesthetics was my next consideration. I could have just built a box, but it wouldn’t have been nearly as fun for me. Aesthetics might not be important to you. They are for me. I also like gadgets and that played into at least one purchase for this build. I used PC Part Picker to put together my component list. I found it invaluable during the process and I’d recommend it to everyone. One caveat is that I wouldn’t trust the compatibility aspects. It does a pretty good job of not steering you wrong, but do your own research. The rest of it isn’t really sexy. I started out with what appealed to me and then I made changes and additions as I dived deep into researching each component and interaction I could find. The resources I used are innumerable. I used reviews, product descriptions, forum posts (praises and problems), et al. to assist me. I also asked friends into gaming what they thought about my component list. And when I got near the end I posted my list to the Reddit /r/buildapc forum. I cannot stress the value of extra sets of eyeballs and first hand experiences. Some of the resources I used: PC Part Picker Tom’s Hardware bit-tech Reddit Purchase PC Part Picker favors certain vendors. You should look at others too. In my case I found their favorites to be the best. My priorities were out-the-door price and shipping time. I knew that once I started getting parts I’d want to start building. Luckily, I timed it well and everything arrived within the span of a few days. Here are my opinions on the vendors I ended up using in alphabetical order. Amazon.com is a good, reliable choice. They have excellent customer service in my experience, and I knew I wouldn’t have trouble with them. However, shipping time is often a problem when you use their free shipping unless you order expensive items (I’ve found items over $100 ship quickly). Ultimately though, price wasn’t always the best and their collection of sales tax in my state turned me off them. I did purchase my case from them. I ordered the mouse as well, but I cancelled after it was stuck four days in a “shipping soon” state. I purchased the mouse locally. Best Buy is not my favorite place to do business. There’s a lot of history with poor, uninterested sales representatives and they used to have a lot of bad anti-consumer policies. That’s a lot better now, but the bad taste is still in my mouth. I ended up purchasing the accessories from them including mouse (locally) and headphones. NCIX is a company that I’ve never heard of before. It popped up as a recommendation for my CPU cooler on PC Part Picker. I didn’t do a lot of research on the company, because their policy on you buying insurance for your orders turned me off. That policy makes it clear to me that the company finds me responsible for the shipment once it leaves their dock. That’s not right, and may run afoul of state laws. Regardless they shipped my CPU cooler quickly and I didn’t have a problem. NewEgg.com is a well known company. I had never done business with them, but I’m glad I did. They shipped quickly and provided good visibility over everything. The prices were also the best in most cases. My main complaint is that they have a lot of exchange only return policies on components. To their credit those policies are listed in the cart underneath each item. The visibility tells me that they’re not playing any shenanigans and made me comfortable dealing with that risk. The vast majority of what I ordered came from them. Coming Next In the next part I’ll tackle my build experience.

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  • Building an OpenStack Cloud for Solaris Engineering, Part 1

    - by Dave Miner
    One of the signature features of the recently-released Solaris 11.2 is the OpenStack cloud computing platform.  Over on the Solaris OpenStack blog the development team is publishing lots of details about our version of OpenStack Havana as well as some tips on specific features, and I highly recommend reading those to get a feel for how we've leveraged Solaris's features to build a top-notch cloud platform.  In this and some subsequent posts I'm going to look at it from a different perspective, which is that of the enterprise administrator deploying an OpenStack cloud.  But this won't be just a theoretical perspective: I've spent the past several months putting together a deployment of OpenStack for use by the Solaris engineering organization, and now that it's in production we'll share how we built it and what we've learned so far.In the Solaris engineering organization we've long had dedicated lab systems dispersed among our various sites and a home-grown reservation tool for developers to reserve those systems; various teams also have private systems for specific testing purposes.  But as a developer, it can still be difficult to find systems you need, especially since most Solaris changes require testing on both SPARC and x86 systems before they can be integrated.  We've added virtual resources over the years as well in the form of LDOMs and zones (both traditional non-global zones and the new kernel zones).  Fundamentally, though, these were all still deployed in the same model: our overworked lab administrators set up pre-configured resources and we then reserve them.  Sounds like pretty much every traditional IT shop, right?  Which means that there's a lot of opportunity for efficiencies from greater use of virtualization and the self-service style of cloud computing.  As we were well into development of OpenStack on Solaris, I was recruited to figure out how we could deploy it to both provide more (and more efficient) development and test resources for the organization as well as a test environment for Solaris OpenStack.At this point, let's acknowledge one fact: deploying OpenStack is hard.  It's a very complex piece of software that makes use of sophisticated networking features and runs as a ton of service daemons with myriad configuration files.  The web UI, Horizon, doesn't often do a good job of providing detailed errors.  Even the command-line clients are not as transparent as you'd like, though at least you can turn on verbose and debug messaging and often get some clues as to what to look for, though it helps if you're good at reading JSON structure dumps.  I'd already learned all of this in doing a single-system Grizzly-on-Linux deployment for the development team to reference when they were getting started so I at least came to this job with some appreciation for what I was taking on.  The good news is that both we and the community have done a lot to make deployment much easier in the last year; probably the easiest approach is to download the OpenStack Unified Archive from OTN to get your hands on a single-system demonstration environment.  I highly recommend getting started with something like it to get some understanding of OpenStack before you embark on a more complex deployment.  For some situations, it may in fact be all you ever need.  If so, you don't need to read the rest of this series of posts!In the Solaris engineering case, we need a lot more horsepower than a single-system cloud can provide.  We need to support both SPARC and x86 VM's, and we have hundreds of developers so we want to be able to scale to support thousands of VM's, though we're going to build to that scale over time, not immediately.  We also want to be able to test both Solaris 11 updates and a release such as Solaris 12 that's under development so that we can work out any upgrade issues before release.  One thing we don't have is a requirement for extremely high availability, at least at this point.  We surely don't want a lot of down time, but we can tolerate scheduled outages and brief (as in an hour or so) unscheduled ones.  Thus I didn't need to spend effort on trying to get high availability everywhere.The diagram below shows our initial deployment design.  We're using six systems, most of which are x86 because we had more of those immediately available.  All of those systems reside on a management VLAN and are connected with a two-way link aggregation of 1 Gb links (we don't yet have 10 Gb switching infrastructure in place, but we'll get there).  A separate VLAN provides "public" (as in connected to the rest of Oracle's internal network) addresses, while we use VxLANs for the tenant networks. One system is more or less the control node, providing the MySQL database, RabbitMQ, Keystone, and the Nova API and scheduler as well as the Horizon console.  We're curious how this will perform and I anticipate eventually splitting at least the database off to another node to help simplify upgrades, but at our present scale this works.I had a couple of systems with lots of disk space, one of which was already configured as the Automated Installation server for the lab, so it's just providing the Glance image repository for OpenStack.  The other node with lots of disks provides Cinder block storage service; we also have a ZFS Storage Appliance that will help back-end Cinder in the near future, I just haven't had time to get it configured in yet.There's a separate system for Neutron, which is our Elastic Virtual Switch controller and handles the routing and NAT for the guests.  We don't have any need for firewalling in this deployment so we're not doing so.  We presently have only two tenants defined, one for the Solaris organization that's funding this cloud, and a separate tenant for other Oracle organizations that would like to try out OpenStack on Solaris.  Each tenant has one VxLAN defined initially, but we can of course add more.  Right now we have just a single /24 network for the floating IP's, once we get demand up to where we need more then we'll add them.Finally, we have started with just two compute nodes; one is an x86 system, the other is an LDOM on a SPARC T5-2.  We'll be adding more when demand reaches the level where we need them, but as we're still ramping up the user base it's less work to manage fewer nodes until then.My next post will delve into the details of building this OpenStack cloud's infrastructure, including how we're using various Solaris features such as Automated Installation, IPS packaging, SMF, and Puppet to deploy and manage the nodes.  After that we'll get into the specifics of configuring and running OpenStack itself.

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  • Webcast Q&A: Qualcomm Provides a Seamless Experience for Customers with Oracle WebCenter

    - by kellsey.ruppel
    Last Thursday we had the second webcast in our WebCenter in Action webcast series, "Qualcomm Provides a Seamless Experience for Customers with Oracle WebCenter, where customer Michael Chander from Qualcomm and Vince Casarez & Gourav Goyal from Oracle Partner Keste shared how Oracle WebCenter is powering Qualcomm’s externally facing website and providing a seamless experience for their customers. In case you missed it, here's a recap of the Q&A.   Mike Chandler, Qualcomm Q: Did you run into any issues when integrating all of the different applications together?A: Definitely, our main challenges were in the area of user provisioning and security propagation, all the standard stuff you might expect when hooking up SSO for authentication and authorization. In addition, we spent several iterations getting the UI’s in sync. While everyone was given the same digital material to build too, each team interpreted and implemented it their own way. Initially as a user navigated, if you were looking for it, you could slight variations in color or font or width , stuff like that. So we had to pull all the developers responsible for the UI together and get pixel level agreement on a lot of things so we could ensure seamless transitions across applications. Q: What has been the biggest benefit your end users have seen?A: Wow, there have been several. An SSO enabled environment was huge a win for our users. The portal application that this replaced had not really been invested in by the business. With this project, we had full business participation and backing, and it really showed in some key areas like the shopping experience. For example, while ordering in the previous site, the items did not have any pictures or really usable descriptions. A tremendous amount of work was done to try and make the site more intuitive and user friendly. Site performance has also drastically improved thanks to new hardware, improved database design, and of course the fact that ADF has made great strides in runtime performance. Q: Was there any resistance internally when implementing the solution? If so, how did you overcome that?A: Within a large company, I’m sure there is always going to be competition for large projects, as there was here. Once we got through the technical analysis and settled on the technology choices, it was actually no resistance to implementing the solution. This project was fully driven by the business with the aim of long term growth. I can confidently say that the fact that this project was given the utmost importance by both the business and IT really help put down any resistance that you would typically see while implementing a new solution. Q: Given the performance, what do you estimate to be the top end capacity of the system? A:I think our top end capacity is really only limited by our hardware. I’m comfortable saying we could grow 10x on our current hardware, both in terms of transactions and users. We can easily spin up new JVM instances if needed. We already use less JVM’s than we had planned. In addition, ADF is doing a very good job with his connection pooling and application module pooling, so we see a very good ratio of users connected to the systems vs db connections, without impacting performace. Q: What's the overview or summary of feedback from the users interacting with the site?A: Feedback has been overwhelmingly positive from both the business and our customers. They’re very happy with the new SSO environment , the new LAF, and the performance of the site. Of course, it’s not all roses. No matter what, there are always going to be people that don’t like the layout or the color scheme, etc. By and large though, customers are happy and the business is happy. Q: Can you describe the impressions about the site before and after the project within Qualcomm?A: Before the project, the site worked and people were using it, but most people were not happy with it. It was slow and tended to be a bit tempermental, for example a user would perform a transaction and the system would throw and unexpected error. The user could back up and retry the steps and things would work fine, so why didn’t work the first time?. From a UI perspective, we’d hear comments like it looked like it was built by a high school student.  Vince Casarez & Gourav Goyal, Keste Q: Did you run into any obstacles when implementing the solution?A: It's interesting some people call them "obstacles" on this project we just called them "dependencies".  There were both technical and business related dependencies that we had to work out. Mike points out the SSO dependencies and the coordination and synchronization between the teams to have a seamless login experience and a seamless end user experience.  There was also a set of dependencies on the User Acceptance testing to make sure that everyone understood the use cases for how the system would be used.  With a branching into a new market and trying to match a simple user experience as many consumer sites have today, there was always a tendency for the team members to provide their suggestions on how things could be simpler.  But with all the work up front on the user design and getting the business driving this set of experiences, this minimized the downstream suggestions that tend to distract a team.  In this case, all the work up front allowed us to enumerate the "dependencies" and keep the distractions to a minimum. Q: Was there a lot of custom work that needed to be done for this particular solution?A: The focus for this particular solution was really on the custom processes. The interesting thing is that with the data flows and the integration with applications, there are some pre-built integrations, but realistically for the process flow, we had to build those. The framework and tooling we used made things easier so we didn’t have to implement core functionality, like transitioning from screen to screen or from flow to flow. The design feature of Task Flows really helped speed the development and keep the component infrastructure in line with the dynamic processes.  Task flows and other elements like Skins are core to the infrastructure or technology stack of Oracle. This then allowed the team to center the project focus around the business flows and use cases to meet the core requirements and keep the project on time. Q: What do you think were the keys to success for rolling out WebCenter?A:  The 5 main keys to success were: 1) Sponsorship from the whole organization around this project from senior executive agreement, business owners driving functionality, and IT development alignment; 2) Upfront design planning and use case definition to clearly define the project scope and requirements; 3) Focussed development and project management aligned with the top level goals and drivers; 4) User acceptance and usability testing along the way to identify potential issues and direct resolution of the issues;  and 5) Constant prioritization of the issues for development to fix by the business.  It also helps to have great team chemistry and really smart people working on the project. If you missed the webcast, be sure to catch the replay to see a live demonstration of WebCenter in action!  Qualcomm Provides a Seamless Experience for Customers with Oracle WebCenter from Oracle WebCenter

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  • DBA Best Practices - A Blog Series: Episode 1 - Backups

    - by Argenis
      This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups have completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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  • SOA Suite Integration: Part 3: Loading files

    - by Anthony Shorten
    One of the most common scenarios in SOA Integration is the loading of a file into the product from an external source. In Oracle SOA Suite there is a File Adapter that can process many file types into your BPEL process. For this example I will use the File Adapter to load a file of user and emails to update the user object within the Oracle Utilities Application Framework. Remember you can repeat this process with other objects and other file types. Again I am illustrating the ease of integration. The first thing is to create an empty BPEL process that will hold our flow. In Oracle JDeveloper this can be achieved by specifying the Define Service Later template (as other templates have predefined inputs and outputs and in this case we want to specify those). So I will create simpleFileLoad process to house our process. You will start with an empty canvas so you need to first specify the load part of the process using the File Adapter. Select the File Adapter from the Component Palette under BPEL Services and drag and drop it to the left side Partner Links (left is input). You name the Service. In this case I chose LoadFile. Press Next. We will define the interface as part of the wizard so select Define from operation and schema (specified later). Press Next. We are going to choose Read File to denote that we will read the file and specify the default Operation Name as Read. Press Next. The next step is to tell the Adapter the location of the files, how to process them and what to do with them after they have been processed. I am using hardcoded locations in this example but you can have logical locations as well. Press Next. I am now going to tell the adapter how to recognize the files I want to load. In my case I am using CSV files and more importantly I am tell the adapter to run the process for each record in the file it encounters. Press Next. Now, I tell the adapter how often I want to poll for the files. I have taken the defaults. Press Next. At this stage I have no explanation of the format of the input. So I am going to invoke the Native Format Wizard which will guide me through the process of creating the file input format. Clicking the purple cog icon will start the wizard. After an introduction screen (not shown), you specify the format of the input file. The File Adapter supports multiple format types. For this example, I will use Delimited as I am going to load a CSV file. Press Next. The best way for the wizard to work is with a sample. I have a sample file and the wizard will ask how much of the file to use as a template. I will use the defaults. Note: If you are using a language that has other languages other than US-ASCII, it is at this point you specify the character set to use.  Press Next. The sample contains multiple instances of a single record type. The wizard supports complex types as well. We will use the appropriate setting for our file. Press Next. You have to specify the file element and the record element. This will be used by the input wizard to translate the CSV data into an XML structure (this will make sense later). I am using LoadUsers as my file delimiter (root element) and User Record as my record root element. Press Next. As the file is CSV the delimiter is "," so I will also specify that the End Of Line (EOL) indicator indicates the end of a record. Press Next. Up until this point your have not given the columns their names. In my case my sample includes the column names in the first record. This is not always the case but you can specify the names and formats of columns in this dialog (not shown). Press Next. The wizard now generates the schema for the input file. You can specify a name for the schema. I have used userupdate.xsd. We want to verify the schema so press Test. You can test the schema by specifying an input sample. and pressing the green play button. You will see the delimiters you specified earlier for the file and the records. Press Ok to continue. A confirmation screen will be displayed showing you the location of the schema in your project. Press Finish to return to the File Adapter configuration. You will now see the schema and elements prepopulated from the wizard. Press Next. The File Adapter configuration is now complete. Press Finish. Now you need to receive the input from the LoadFile component so we need to place a Receive node in the BPEL process by drag and dropping the Receive component from the Component Palette under BPEL Constructs onto the BPEL process. We link the receive process with the LoadFile component by dragging the left most connect node of the Receive node to the LoadFile component. Once the link is established you need to name the Receive node appropriately and as in the post of the last part of this series you need to generate input variables for the BPEL process to hold the input records in. You need to now add the product Web Service. The process is the same as described in the post of the last part of this series. You drop the Web Service BPEL Service onto the right side of the process and fill in the details of the WSDL URL . You also have to add an Invoke node to call the service and generate the input and outputs variables for the call in the Invoke node. Now, to get the inputs from File to the service. You have to use a Transform (you can use an Assign action but a Transform action is more flexible). You drag and drop the Transform component from the Component Palette under Oracle Extensions and place it between the Receive and Invoke nodes. We name the Transform Node, Mapper File and associate the source of the mapping the schema from the Receive node and the output will be the input variable from the Invoke node. We now build the transform. We first map the user and email attributes by drag and drop the elements from the left to the right. The reason we needed to use the transform is that we will be telling the AS-User service that we want to issue an update action. Remember when we registered the service we actually used Read as the default. If we do not otherwise inform the service to use the Update action it will use the Read action instead (which is not desired). To specify the update action you need to click on the transactionType node on the right and select Set Text to set the action. You need to specify the transactionType of UPD (for update). The mapping is now complete. The final BPEL process is ready for deployment. You then deploy the BPEL process to the server and to test the service by simply dropping a file, in the same pattern/name as you specified, in the directory you specified in the File Adapter. You will see each record as a separate instance entry in the Fusion Middleware Control console. You can now load files into the product. You can repeat this process for each type of file to process. While this was a simple example it illustrates the method of loading data can be achieved using SOA Suite in conjunction with our products.

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  • Model View Control Issue: Null Pointer Initialization Question [closed]

    - by David Dimalanta
    Good morning again. This is David. Please, I need an urgent help regarding control model view where I making a code that uniquely separating into groups: An Activity Java Class to Display the Interface A View and Function Java Class for Drawing Cards and Display it on the Activity Class The problem is that the result returns a Null Pointer Exception. I have initialize for the ID for Text View and Image View. Under this class "draw_deck.java". Please help me. Here's my code for draw_deck.java: package com.bodapps.inbetween.model; import android.content.Context; import android.view.View; import android.widget.ImageView; import android.widget.TextView; import com.bodapps.inbetween.R; public class draw_deck extends View { public TextView count_label; public ImageView draw_card; private int count; public draw_deck(Context context) { super(context); // TODO Auto-generated constructor stub //I have initialized two widgets for ID. I still don't get it why I got forced closed by Null Pointer Exception thing. draw_card = (ImageView) findViewById(R.id.IV_Draw_Card); count_label = (TextView) findViewById(R.id.Text_View_Count_Card); } public void draw(int s, int c, String strSuit, String strValue, Pile pile, Context context) { //super(context); //Just printing the card drawn from pile int suit, value = 1; draw_card = (ImageView) findViewById(R.id.IV_Draw_Card); count_label = (TextView) findViewById(R.id.Text_View_Count_Card); Card card; if(!pile.isEmpty()) //Setting it to IF statement displays the card one by one. { card = pile.drawFromPile(); //Need to check first if card is null. if (card != null) { //draws an extra if (card != null) { //Get suit of card to print out. suit = card.getSuit(); switch (suit) { case CardInfo.DIAMOND: strSuit = "DIAMOND"; s=0; break; case CardInfo.HEART: strSuit = "HEART"; s=1; break; case CardInfo.SPADE: strSuit = "SPADE"; s=2; break; case CardInfo.CLUB: strSuit = "CLUB"; s=3; break; } //Get value of card to print out. value = card.getValue(); switch (value) { case CardInfo.ACE: strValue = "ACE"; c=0; break; case CardInfo.TWO: c=1; break; case CardInfo.THREE: strValue = "THREE"; c=2; break; case CardInfo.FOUR: strValue = "FOUR"; c=3; break; case CardInfo.FIVE: strValue = "FIVE"; c=4; break; case CardInfo.SIX: strValue = "SIX"; c=4; break; case CardInfo.SEVEN: strValue = "SEVEN"; c=4; break; case CardInfo.EIGHT: strValue = "EIGHT"; c=4; break; case CardInfo.NINE: strValue = "NINE"; c=4; break; case CardInfo.TEN: strValue = "TEN"; c=4; break; case CardInfo.JACK: strValue = "JACK"; c=4; break; case CardInfo.QUEEN: strValue = "QUEEN"; c=4; break; case CardInfo.KING: strValue = "KING"; c=4; break; } } } }// //Below two lines of code, this is where issued the Null Pointer Exception. draw_card.setImageResource(deck[s][c]); count_label.setText(new StringBuilder(strValue).append(" of ").append(strSuit).append(String.valueOf(" " + count++)).toString()); } //Choice of Suits in a Deck public Integer[][] deck = { //Array Group 1 is [0][0] (No. of Cards: 4 - DIAMOND) { R.drawable.card_dummy_1, R.drawable.card_dummy_2, R.drawable.card_dummy_4, R.drawable.card_dummy_5, R.drawable.card_dummy_3 }, //Array Group 2 is [1][0] (No. of Cards: 4 - HEART) { R.drawable.card_dummy_1, R.drawable.card_dummy_2, R.drawable.card_dummy_4, R.drawable.card_dummy_5, R.drawable.card_dummy_3 }, //Array Group 3 is [2][0] (No. of Cards: 4 - SPADE) { R.drawable.card_dummy_1, R.drawable.card_dummy_2, R.drawable.card_dummy_4, R.drawable.card_dummy_5, R.drawable.card_dummy_3 }, //Array Group 4 is [3][0] (No. of Cards: 4 - CLUB) { R.drawable.card_dummy_1, R.drawable.card_dummy_2, R.drawable.card_dummy_4, R.drawable.card_dummy_5, R.drawable.card_dummy_3 }, }; } And this one of the activity class, Player_Mode_2.java: package com.bodapps.inbetween; import java.util.Random; import android.app.Activity; import android.app.Dialog; import android.content.Context; import android.os.Bundle; import android.view.View; import android.view.View.OnClickListener; import android.widget.Button; import android.widget.EditText; import android.widget.ImageView; import android.widget.TextView; import android.widget.Toast; import com.bodapps.inbetween.model.Card; import com.bodapps.inbetween.model.Pile; import com.bodapps.inbetween.model.draw_deck; /* * * Public class for Two-Player mode. * */ public class Player_Mode_2 extends Activity { //Image Views private ImageView draw_card; private ImageView player_1; private ImageView player_2; private ImageView icon; //Buttons private Button set_deck; //Edit Texts private EditText enter_no_of_decks; //text Views private TextView count_label; //Integer Data Types private int no_of_cards, count; private int card_multiplier; //Contexts final Context context = this; //Pile Model public Pile pile; //Card Model public Card card; //create View @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.play_2_player_mode); //-----[ Search for Views ]----- //Initialize for Image View draw_card = (ImageView) findViewById(R.id.IV_Draw_Card); player_1 = (ImageView) findViewById(R.id.IV_Player_1_Card); player_2 = (ImageView) findViewById(R.id.IV_Player_2_Card); //Initialize for Text view or Label count_label = (TextView) findViewById(R.id.Text_View_Count_Card); //-----[ Adding Values ]----- //Integer Values count = 0; no_of_cards = 0; //-----[ Adding Dialog ]----- //Initializing Dialog final Dialog deck_dialog = new Dialog(context); deck_dialog.setContentView(R.layout.dialog); deck_dialog.setTitle("Deck Dialog"); //-----[ Initializing Views for Dialog's Contents ]----- //Initialize for Edit Text enter_no_of_decks = (EditText) deck_dialog.findViewById(R.id.Edit_Text_Set_Number_of_Decks); //Initialize for Button set_deck = (Button) deck_dialog.findViewById(R.id.Button_Deck); //-----[ Setting onClickListener() ]----- //Set Event Listener for Image view draw_card.setOnClickListener(new Draw_Card_Model()); //Set Event Listener for Setting the Deck set_deck.setOnClickListener(new OnClickListener() { public void onClick(View v) { if(card_multiplier <= 8) { //Use "Integer.parseInt()" method to instantly convert from String to int value. card_multiplier = Integer.parseInt(enter_no_of_decks.getText().toString()); //Shuffling cards... pile = new Pile(card_multiplier); //Multiply no. of decks //Dismiss or close the dialog. deck_dialog.dismiss(); } else { Toast.makeText(getApplicationContext(), "Please choose a number from 1 to 8.", Toast.LENGTH_SHORT).show(); } } }); //Show dialog. deck_dialog.show(); } //Shuffling the Array public void Shuffle_Cards(Integer[][] Shuffle_Deck) { Random random = new Random(); for(int i = Shuffle_Deck[no_of_cards].length - 1; i >=0; i--) { int Index = random.nextInt(i + 1); //Simple Swapping Integer swap = Shuffle_Deck[card_multiplier-1][Index]; Shuffle_Deck[card_multiplier-1][Index] = Shuffle_Deck[card_multiplier-1][i]; Shuffle_Deck[card_multiplier-1][i] = swap; } } //Private Class for Random Card Draw private class Draw_Card_Model implements OnClickListener { public void onClick(View v) { //Just printing the card drawn from pile int suit = 0, value = 0; String strSuit = "", strValue = ""; draw_deck draw = new draw_deck(context); //This line is where issued the Null Pointer Exception. if (count == card_multiplier*52) { // A message shows up when all cards are draw out. Toast.makeText(getApplicationContext(), "All cards have been used up.", Toast.LENGTH_SHORT).show(); draw_card.setEnabled(false); } else { draw.draw(suit, value, strSuit, strValue, pile, context); count_label.setText(count); //This is where I got force closed error, although "int count" have initialized the number. This was supposed to accept in the setText() method. count++; } } } } Take note that the issues on Null Pointer Exception is the Image View and the Edit Text. I got to test it. Thanks. If you have any info about my question, let me know it frankly.

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  • OData &ndash; The easiest service I can create: now with updates

    - by Jon Dalberg
    The other day I created a simple NastyWord service exposed via OData. It was read-only and used an in-memory backing store for the words. Today I’ll modify it to use a file instead of a list and I’ll accept new nasty words by implementing IUpdatable directly. The first thing to do is enable the service to accept new entries. This is done at configuration time by adding the “WriteAppend” access rule: 1: public class NastyWords : DataService<NastyWordsDataSource> 2: { 3: // This method is called only once to initialize service-wide policies. 4: public static void InitializeService(DataServiceConfiguration config) 5: { 6: config.SetEntitySetAccessRule("*", EntitySetRights.AllRead | EntitySetRights.WriteAppend); 7: config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; 8: } 9: }   Next I placed a file, NastyWords.txt, in the “App_Data” folder and added a few *choice* words to start. This required one simple change to our NastyWordDataSource.cs file: 1: public NastyWordsDataSource() 2: { 3: UpdateFromSource(); 4: } 5:   6: private void UpdateFromSource() 7: { 8: var words = File.ReadAllLines(pathToFile); 9: NastyWords = (from w in words 10: select new NastyWord { Word = w }).AsQueryable(); 11: }   Nothing too shocking here, just reading each line from the NastyWords.txt file and exposing them. Next, I implemented IUpdatable which comes with a boat-load of methods. We don’t need all of them for now since we are only concerned with allowing new values. Here are the methods we must implement, all the others throw a NotImplementedException: 1: public object CreateResource(string containerName, string fullTypeName) 2: { 3: var nastyWord = new NastyWord(); 4: pendingUpdates.Add(nastyWord); 5: return nastyWord; 6: } 7:   8: public object ResolveResource(object resource) 9: { 10: return resource; 11: } 12:   13: public void SaveChanges() 14: { 15: var intersect = (from w in pendingUpdates 16: select w.Word).Intersect(from n in NastyWords 17: select n.Word); 18:   19: if (intersect.Count() > 0) 20: throw new DataServiceException(500, "duplicate entry"); 21:   22: var lines = from w in pendingUpdates 23: select w.Word; 24:   25: File.AppendAllLines(pathToFile, 26: lines, 27: Encoding.UTF8); 28:   29: pendingUpdates.Clear(); 30:   31: UpdateFromSource(); 32: } 33:   34: public void SetValue(object targetResource, string propertyName, object propertyValue) 35: { 36: targetResource.GetType().GetProperty(propertyName).SetValue(targetResource, propertyValue, null); 37: }   I use a simple list to contain the pending updates and only commit them when the “SaveChanges” method is called. Here’s the order these methods are called in our service during an insert: CreateResource – here we just instantiate a new NastyWord and stick a reference to it in our pending updates list. SetValue – this is where the “Word” property of the NastyWord instance is set. SaveChanges – get the list of pending updates, barfing on duplicates, write them to the file and clear our pending list. ResolveResource – the newly created resource will be returned directly here since we aren’t dealing with “handles” to objects but the actual objects themselves. Not too bad, eh? I didn’t find this documented anywhere but a little bit of digging in the OData spec and use of Fiddler made it pretty easy to figure out. Here is some client code which would add a new nasty word: 1: static void Main(string[] args) 2: { 3: var svc = new ServiceReference1.NastyWordsDataSource(new Uri("http://localhost.:60921/NastyWords.svc")); 4: svc.AddToNastyWords(new ServiceReference1.NastyWord() { Word = "shat" }); 5:   6: svc.SaveChanges(); 7: }   Here’s all of the code so far for to implement the service: 1: using System; 2: using System.Collections.Generic; 3: using System.Data.Services; 4: using System.Data.Services.Common; 5: using System.Linq; 6: using System.ServiceModel.Web; 7: using System.Web; 8: using System.IO; 9: using System.Text; 10:   11: namespace ONasty 12: { 13: [DataServiceKey("Word")] 14: public class NastyWord 15: { 16: public string Word { get; set; } 17: } 18:   19: public class NastyWordsDataSource : IUpdatable 20: { 21: private List<NastyWord> pendingUpdates = new List<NastyWord>(); 22: private string pathToFile = @"path to your\App_Data\NastyWords.txt"; 23:   24: public NastyWordsDataSource() 25: { 26: UpdateFromSource(); 27: } 28:   29: private void UpdateFromSource() 30: { 31: var words = File.ReadAllLines(pathToFile); 32: NastyWords = (from w in words 33: select new NastyWord { Word = w }).AsQueryable(); 34: } 35:   36: public IQueryable<NastyWord> NastyWords { get; private set; } 37:   38: public void AddReferenceToCollection(object targetResource, string propertyName, object resourceToBeAdded) 39: { 40: throw new NotImplementedException(); 41: } 42:   43: public void ClearChanges() 44: { 45: pendingUpdates.Clear(); 46: } 47:   48: public object CreateResource(string containerName, string fullTypeName) 49: { 50: var nastyWord = new NastyWord(); 51: pendingUpdates.Add(nastyWord); 52: return nastyWord; 53: } 54:   55: public void DeleteResource(object targetResource) 56: { 57: throw new NotImplementedException(); 58: } 59:   60: public object GetResource(IQueryable query, string fullTypeName) 61: { 62: throw new NotImplementedException(); 63: } 64:   65: public object GetValue(object targetResource, string propertyName) 66: { 67: throw new NotImplementedException(); 68: } 69:   70: public void RemoveReferenceFromCollection(object targetResource, string propertyName, object resourceToBeRemoved) 71: { 72: throw new NotImplementedException(); 73: } 74:   75: public object ResetResource(object resource) 76: { 77: throw new NotImplementedException(); 78: } 79:   80: public object ResolveResource(object resource) 81: { 82: return resource; 83: } 84:   85: public void SaveChanges() 86: { 87: var intersect = (from w in pendingUpdates 88: select w.Word).Intersect(from n in NastyWords 89: select n.Word); 90:   91: if (intersect.Count() > 0) 92: throw new DataServiceException(500, "duplicate entry"); 93:   94: var lines = from w in pendingUpdates 95: select w.Word; 96:   97: File.AppendAllLines(pathToFile, 98: lines, 99: Encoding.UTF8); 100:   101: pendingUpdates.Clear(); 102:   103: UpdateFromSource(); 104: } 105:   106: public void SetReference(object targetResource, string propertyName, object propertyValue) 107: { 108: throw new NotImplementedException(); 109: } 110:   111: public void SetValue(object targetResource, string propertyName, object propertyValue) 112: { 113: targetResource.GetType().GetProperty(propertyName).SetValue(targetResource, propertyValue, null); 114: } 115: } 116:   117: public class NastyWords : DataService<NastyWordsDataSource> 118: { 119: // This method is called only once to initialize service-wide policies. 120: public static void InitializeService(DataServiceConfiguration config) 121: { 122: config.SetEntitySetAccessRule("*", EntitySetRights.AllRead | EntitySetRights.WriteAppend); 123: config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; 124: } 125: } 126: } Next time we’ll allow removing nasty words. Enjoy!

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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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