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• Formal Languages, Inductive Proofs &amp; Regular Expressions

  - by MarkPearl

  So I am slogging away at my UNISA stuff. I have just finished doing the initial once non stop read through the first 11 chapters of my COS 201 Textbook - “Introduction to Computer Theory 2nd Edition” by Daniel Cohen. It has been an interesting couple of days, with familiar concepts coming up as well as some new territory. In this posting I am going to cover the first couple of chapters of the book. Let start with Formal Languages… What exactly is a formal language? Pretty much a no duh question for me but still a good one to ask – a formal language is a language that is defined in a precise mathematical way. Does that mean that the English language is a formal language? I would say no – and my main motivation for this is that one can have an English sentence that is correct grammatically that is also ambiguous. For example the ambiguous sentence: "I once shot an elephant in my pyjamas.” For this and possibly many other reasons that I am unaware of, English is termed a “Natural Language”. So why the importance of formal languages in computer science? Again a no duh question in my mind… If we want computers to be effective and useful tools then we need them to be able to evaluate a series of commands in some form of language that when interpreted by the device no confusion will exist as to what we were requesting. Imagine the mayhem that would exist if a computer misinterpreted a command to print a document and instead decided to delete it. So what is a Formal Language made up of… For my study purposes a language is made up of a finite alphabet. For a formal language to exist there needs to be a specification on the language that will describe whether a string of characters has membership in the language or not. There are two basic ways to do this: By a “machine” that will recognize strings of the language (e.g. Finite Automata). By a rule that describes how strings of a language can be formed (e.g. Regular Expressions). When we use the phrase “string of characters”, we can also be referring to a “word”. What is an Inductive Proof? So I am not to far into my textbook and of course it starts referring to proofs and different types. I have had to go through several different approaches of proofs in the past, but I can never remember their formal names , so when I saw “inductive proof” I thought to myself – what the heck is that? Google to the rescue… An inductive proof is like a normal proof but it employs a neat trick which allows you to prove a statement about an arbitrary number n by first proving it is true when n is 1 and then assuming it is true for n=k and showing it is true for n=k+1. The idea is that if you want to show that someone can climb to the nth floor of a fire escape, you need only show that you can climb the ladder up to the fire escape (n=1) and then show that you know how to climb the stairs from any level of the fire escape (n=k) to the next level (n=k+1). Does this sound like a form of recursion? No surprise then that in the same chapter they deal with recursive definitions. An example of a recursive definition for the language EVEN would the 3 rules below: 2 is in EVEN If x is in EVEN then so is x+2 The only elements in the set EVEN are those that be produced by the rules above. Nothing to exciting… So if a definition for a language is done recursively, then it makes sense that the language can be proved using induction. Regular Expressions So I am wondering to myself what use is this all – in fact – I find this the biggest challenge to any university material is that it is quite hard to find the immediate practical applications of some theory in real life stuff. How great was my joy when I suddenly saw the word regular expression being introduced. I had been introduced to regular expressions on Stack Overflow where I was trying to recognize if some text measurement put in by a user was in a valid form or not. For instance, the imperial system of measurement where you have feet and inches can be represented in so many different ways. I had eventually turned to regular expressions as an easy way to check if my parser could correctly parse the text or not and convert it to a normalize measurement. So some rules about languages and regular expressions… Any finite language can be represented by at least one if not more regular expressions A regular expressions is almost a rule syntax for expressing how regular languages can be formed regular expressions are cool For a regular expression to be valid for a language it must be able to generate all the words in the language and no other words. This is important. It doesn’t help me if my regular expression parses 100% of my measurement texts but also lets one or two invalid texts to pass as well. Okay, so this posting jumps around a bit – but introduces some very basic fundamentals for the subject which will be built on in later postings… Time to go and do some practical examples now…

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• PHP OCI8 and Oracle 11g DRCP Connection Pooling in Pictures

  - by christopher.jones

  Here is a screen shot from a PHP OCI8 connection pooling demo that I like to run. It graphically shows how little database host memory is needed when using DRCP connection pooling with Oracle Database 11g. Migrating to DRCP can be as simple as starting the pool and changing the connection string in your PHP application. The script that generated the data for this graph was a simple "Parts" query application being run under various simulated user loads. I was running the database on a small Oracle Linux server with just 2G of memory. I used PHP OCI8 1.4. Apache is in pre-fork mode, as needed for PHP. Each graph has time on the horizontal access in arbitrary 'tick' time units. Click the image to see it full sized. Pooled connections Beginning with the top left graph, At tick time 65 I used Apache's 'ab' tool to start 100 concurrent 'users' running the application. These users connected to the database using DRCP: $c = oci_pconnect('phpdemo', 'welcome', 'myhost/orcl:pooled'); A second hundred DRCP users were added to the system at tick 80 and a final hundred users added at tick 100. At about tick 110 I stopped the test and restarted Apache. This closed all the connections. The bottom left graph shows the number of statements being executed by the database per second, with some spikes for background database activity and some variability for this small test. Each extra batch of users adds another 'step' of load to the system. Looking at the top right Server Process graph shows the database server processes doing the query work for each web user. As user load is added, the DRCP server pool increases (in green). The pool is initially at its default size 4 and quickly ramps up to about (I'm guessing) 35. At tick time 100 the pool increases to my configured maximum of 40 processes. Those 40 processes are doing the query work for all 300 web users. When I stopped the test at tick 110, the pooled processes remained open waiting for more users to connect. If I had left the test quiet for the DRCP 'inactivity_timeout' period (300 seconds by default), the pool would have shrunk back to 4 processes. Looking at the bottom right, you can see the amount of memory being consumed by the database. During the initial quiet period about 500M of memory was in use. The absolute number is just an indication of my particular DB configuration. As the number of pooled processes increases, each process needs more memory. You can see the shape of the memory graph echoes the Server Process graph above it. Each of the 300 web users will also need a few kilobytes but this is almost too small to see on the graph. Non-pooled connections Compare the DRCP case with using 'dedicated server' processes. At tick 140 I started 100 web users who did not use pooled connections: $c = oci_pconnect('phpdemo', 'welcome', 'myhost/orcl'); This connection string change is the only difference between the two tests. At ticks 155 and 165 I started two more batches of 100 simulated users each. At about tick 195 I stopped the user load but left Apache running. Apache then gradually returned to its quiescent state, killing idle httpd processes and producing the downward slope at the right of the graphs as the persistent database connection in each Apache process was closed. The Executions per Second graph on the bottom left shows the same step increases as for the earlier DRCP case. The database is handling this load. But look at the number of Server processes on the top right graph. There is now a one-to-one correspondence between Apache/PHP processes and DB server processes. Each PHP processes has one DB server processes dedicated to it. Hence the term 'dedicated server'. The memory required on the database is proportional to all those database server processes started. Almost all my system's memory was consumed. I doubt it would have coped with any more user load. Summary Oracle Database 11g DRCP connection pooling significantly reduces database host memory requirements allow more system memory to be allocated for the SGA and allowing the system to scale to handled thousands of concurrent PHP users. Even for small systems, using DRCP allows more web users to be active. More information about PHP and DRCP can be found in the PHP Scalability and High Availability chapter of The Underground PHP and Oracle Manual.

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• Partition Wise Joins

  - by jean-pierre.dijcks

  Some say they are the holy grail of parallel computing and PWJ is the basis for a shared nothing system and the only join method that is available on a shared nothing system (yes this is oversimplified!). The magic in Oracle is of course that is one of many ways to join data. And yes, this is the old flexibility vs. simplicity discussion all over, so I won't go there... the point is that what you must do in a shared nothing system, you can do in Oracle with the same speed and methods. The Theory A partition wise join is a join between (for simplicity) two tables that are partitioned on the same column with the same partitioning scheme. In shared nothing this is effectively hard partitioning locating data on a specific node / storage combo. In Oracle is is logical partitioning. If you now join the two tables on that partitioned column you can break up the join in smaller joins exactly along the partitions in the data. Since they are partitioned (grouped) into the same buckets, all values required to do the join live in the equivalent bucket on either sides. No need to talk to anyone else, no need to redistribute data to anyone else... in short, the optimal join method for parallel processing of two large data sets. PWJ's in Oracle Since we do not hard partition the data across nodes in Oracle we use the Partitioning option to the database to create the buckets, then set the Degree of Parallelism (or run Auto DOP - see here) and get our PWJs. The main questions always asked are: How many partitions should I create? What should my DOP be? In a shared nothing system the answer is of course, as many partitions as there are nodes which will be your DOP. In Oracle we do want you to look at the workload and concurrency, and once you know that to understand the following rules of thumb. Within Oracle we have more ways of joining of data, so it is important to understand some of the PWJ ideas and what it means if you have an uneven distribution across processes. Assume we have a simple scenario where we partition the data on a hash key resulting in 4 hash partitions (H1 -H4). We have 2 parallel processes that have been tasked with reading these partitions (P1 - P2). The work is evenly divided assuming the partitions are the same size and we can scan this in time t1 as shown below. Now assume that we have changed the system and have a 5th partition but still have our 2 workers P1 and P2. The time it takes is actually 50% more assuming the 5th partition has the same size as the original H1 - H4 partitions. In other words to scan these 5 partitions, the time t2 it takes is not 1/5th more expensive, it is a lot more expensive and some other join plans may now start to look exciting to the optimizer. Just to post the disclaimer, it is not as simple as I state it here, but you get the idea on how much more expensive this plan may now look... Based on this little example there are a few rules of thumb to follow to get the partition wise joins. First, choose a DOP that is a factor of two (2). So always choose something like 2, 4, 8, 16, 32 and so on... Second, choose a number of partitions that is larger or equal to 2* DOP. Third, make sure the number of partitions is divisible through 2 without orphans. This is also known as an even number... Fourth, choose a stable partition count strategy, which is typically hash, which can be a sub partitioning strategy rather than the main strategy (range - hash is a popular one). Fifth, make sure you do this on the join key between the two large tables you want to join (and this should be the obvious one...). Translating this into an example: DOP = 8 (determined based on concurrency or by using Auto DOP with a cap due to concurrency) says that the number of partitions >= 16. Number of hash (sub) partitions = 32, which gives each process four partitions to work on. This number is somewhat arbitrary and depends on your data and system. In this case my main reasoning is that if you get more room on the box you can easily move the DOP for the query to 16 without repartitioning... and of course it makes for no leftovers on the table... And yes, we recommend up-to-date statistics. And before you start complaining, do read this post on a cool way to do stats in 11.

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• Review of ComponentOne Silverlight Controls (Free License Giveaway).

  - by mbcrump

  ComponentOne has several great products that target Silverlight Developers. One of them is their Silverlight Controls and the other is the XAP Optimizer. I decided that I would check out the controls and Xap Optimizer and feature them on my blog. After talking with ComponentOne, they agreed to take part in my Monthly Silverlight giveaway. The details are listed below: ----------------------------------------------------------------------------------------------------------------------------------------------------------- Win a FREE developer’s license of ComponentOne Silverlight Controls + XAP Optimizer! (the winner also gets a license to Silverlight Spy) Random winner will be announced on March 1st, 2011! To be entered into the contest do the following things: Subscribe to my feed. Leave a comment below with a valid email account (I WILL NOT share this info with anyone.) Retweet the following : I just entered to win free #Silverlight controls from @mbcrump and @ComponentOne http://mcrump.me/fTSmB8 ! Don’t change the URL because this will allow me to track the users that Tweet this page. Don’t forget to visit ComponentOne because they made this possible. MichaelCrump.Net provides Silverlight Giveaways every month. You can also see the latest giveaway by bookmarking http://giveaways.michaelcrump.net . ---------------------------------------------------------------------------------------------------------------------------------------------------------- Before we get started with the Silverlight Controls, here is a couple of links to bookmark: The Live Demos of the Silverlight Controls is located here. The XAP Optimizer page is here. One thing that I liked about the help documentation is that you can grab a PDF that only contains documentation for that control. This allows you to get the information you need without going through several hundred pages. You can also download the full documentation from their site.  ComponentOne Silverlight Controls I recently built a hobby project and decided to use ComponentOne Silverlight Controls. The main reason for this is that the controls are heavily documented, they look great and getting help was just a tweet or forum click away. So, the first question that you may ask is, “What is included?” Here is the official list below. I wanted to show several of the controls that I think developers will use the most. 1) ComponentOne’s Image Control – Display animated GIF images on your Silverlight pages as you would in traditional Web apps. Add attractive visuals with minimal effort. 2) HTML Host - Render HTML and arbitrary URI content from within Silverlight. 3) Chart3D - Create 3D surface charts with options for contour levels, zones, a chart legend and more. 4) PDFViewer - View PDF files in Silverlight! That is just a fraction of the controls available. If you want to check out several of them in a “real” application then check out my Silverlight page at http://michaelcrump.info. This brings me to the second part of the giveaway. XAP Optimizer – Is designed to reduce the size of your XAP File. It also includes built-in obfuscation and signing. With my personal project, I decided to use the XAP Optimizer by ComponentOne. It was so easy to use. You basically give it your .XAP file and it provides an output file. If you prefer to prune unused references manually then you can prune your XAP file manually by selecting the option below. I went ahead and added Obfuscation just to try it out and it worked great. You may notice from the screenshot below that I only obfuscated assemblies that I built. The other dlls anyone can grab off the net so we have no reason to obfuscate them. You also have the option to automatically sign your .xap with the SN.exe tool. So how did it turn out? Well, I reduced my XAP size from 2.4 to 1.8 with simply a click of a button. I added obfuscation with a click of a button: Screenshot of no obfuscation on my XAP File   Screenshot of obfuscation on my XAP File with XAP Optimizer.   So, with 2 button clicks, I reduce my XAP file and obfuscated my assembly. What else can you want? Well, they provide a nice HTML report that gives you an optimization summary. So what if you don’t want to launch this tool every time you deploy a Silverlight application? Well the official documentation provided a way to do it in your built event in Visual Studio. Click the Build Events tab on the left side of the Properties window. Enter the following command in the Post-build event command line: $Program Files\ComponentOne\XapOptimizer\XapOptimizer.exe /cmd /p:$(ProjectDir)$(ProjectName).xoproj In the end, this is a great product. I love code that I don’t have to write and utilities that just work. ComponentOne delivers with both the Silverlight Controls and the XAP Optimizer. Don’t forget to leave a comment below in order to win a set of the controls! Subscribe to my feed

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• A C# implementation of the CallStream pattern

  - by Bertrand Le Roy

  Dusan published this interesting post a couple of weeks ago about a novel JavaScript chaining pattern: http://dbj.org/dbj/?p=514 It’s similar to many existing patterns, but the syntax is extraordinarily terse and it provides a new form of friction-free, plugin-less extensibility mechanism. Here’s a JavaScript example from Dusan’s post: CallStream("#container") (find, "div") (attr, "A", 1) (css, "color", "#fff") (logger); The interesting thing here is that the functions that are being passed as the first argument are arbitrary, they don’t need to be declared as plug-ins. Compare that with a rough jQuery equivalent that could look something like this: $.fn.logger = function () { /* ... */ } $("selector") .find("div") .attr("A", 1) .css("color", "#fff") .logger(); There is also the “each” method in jQuery that achieves something similar, but its syntax is a little more verbose. Of course, that this pattern can be expressed so easily in JavaScript owes everything to the extraordinary way functions are treated in that language, something Douglas Crockford called “the very best part of JavaScript”. One of the first things I thought while reading Dusan’s post was how I could adapt that to C#. After all, with Lambdas and delegates, C# also has its first-class functions. And sure enough, it works really really well. After about ten minutes, I was able to write this: CallStreamFactory.CallStream (p => Console.WriteLine("Yay!")) (Dump, DateTime.Now) (DumpFooAndBar, new { Foo = 42, Bar = "the answer" }) (p => Console.ReadKey()); Where the Dump function is: public static void Dump(object options) { Console.WriteLine(options.ToString()); } And DumpFooAndBar is: public static void DumpFooAndBar(dynamic options) { Console.WriteLine("Foo is {0} and bar is {1}.", options.Foo, options.Bar); } So how does this work? Well, it really is very simple. And not. Let’s say it’s not a lot of code, but if you’re like me you might need an Advil after that. First, I defined the signature of the CallStream method as follows: public delegate CallStream CallStream (Action<object> action, object options = null); The delegate define a call stream as something that takes an action (a function of the options) and an optional options object and that returns a delegate of its own type. Tricky, but that actually works, a delegate can return its own type. Then I wrote an implementation of that delegate that calls the action and returns itself: public static CallStream CallStream (Action<object> action, object options = null) { action(options); return CallStream; } Pretty nice, eh? Well, yes and no. What we are doing here is to execute a sequence of actions using an interesting novel syntax. But for this to be actually useful, you’d need to build a more specialized call stream factory that comes with some sort of context (like Dusan did in JavaScript). For example, you could write the following alternate delegate signature that takes a string and returns itself: public delegate StringCallStream StringCallStream(string message); And then write the following call stream (notice the currying): public static StringCallStream CreateDumpCallStream(string dumpPath) { StringCallStream str = null; var dump = File.AppendText(dumpPath); dump.AutoFlush = true; str = s => { dump.WriteLine(s); return str; }; return str; } (I know, I’m not closing that stream; sure; bad, bad Bertrand) Finally, here’s how you use it: CallStreamFactory.CreateDumpCallStream(@".\dump.txt") ("Wow, this really works.") (DateTime.Now.ToLongTimeString()) ("And that is all."); Next step would be to combine this contextual implementation with the one that takes an action parameter and do some really fun stuff. I’m only scratching the surface here. This pattern could reveal itself to be nothing more than a gratuitous mind-bender or there could be applications that we hardly suspect at this point. In any case, it’s a fun new construct. Or is this nothing new? You tell me… Comments are open :)

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• Cant get lm-sensors to load ATI Radeon temp or fan

  - by woody

  New to Linux and having minor issues :/ . I followed this guide initially but did not recieve the proper output and did not show my ATI Radeon HD 5000 temp or fan speed. Then used this guide, same problems exhibited. No issues installing and no errors. I think its not reading i2c for some reason. The proprietary driver is installed and functioning correctly according fglrxinfo. I can use aticonfig commands and view both temp and fan. Any ideas on how to get it working under 'sensors'? When i run 'sudo sensors-detect' this is my ouput # sensors-detect revision 5984 (2011-07-10 21:22:53 +0200) # System: LENOVO IdeaPad Y560 (laptop) # Board: Lenovo KL3 This program will help you determine which kernel modules you need to load to use lm_sensors most effectively. It is generally safe and recommended to accept the default answers to all questions, unless you know what you're doing. Some south bridges, CPUs or memory controllers contain embedded sensors. Do you want to scan for them? This is totally safe. (YES/no): y Silicon Integrated Systems SIS5595... No VIA VT82C686 Integrated Sensors... No VIA VT8231 Integrated Sensors... No AMD K8 thermal sensors... No AMD Family 10h thermal sensors... No AMD Family 11h thermal sensors... No AMD Family 12h and 14h thermal sensors... No AMD Family 15h thermal sensors... No AMD Family 15h power sensors... No Intel digital thermal sensor... Success! (driver `coretemp') Intel AMB FB-DIMM thermal sensor... No VIA C7 thermal sensor... No VIA Nano thermal sensor... No Some Super I/O chips contain embedded sensors. We have to write to standard I/O ports to probe them. This is usually safe. Do you want to scan for Super I/O sensors? (YES/no): y Probing for Super-I/O at 0x2e/0x2f Trying family `National Semiconductor/ITE'... Yes Found unknown chip with ID 0x8502 Probing for Super-I/O at 0x4e/0x4f Trying family `National Semiconductor/ITE'... No Trying family `SMSC'... No Trying family `VIA/Winbond/Nuvoton/Fintek'... No Trying family `ITE'... No Some hardware monitoring chips are accessible through the ISA I/O ports. We have to write to arbitrary I/O ports to probe them. This is usually safe though. Yes, you do have ISA I/O ports even if you do not have any ISA slots! Do you want to scan the ISA I/O ports? (YES/no): y Probing for `National Semiconductor LM78' at 0x290... No Probing for `National Semiconductor LM79' at 0x290... No Probing for `Winbond W83781D' at 0x290... No Probing for `Winbond W83782D' at 0x290... No Lastly, we can probe the I2C/SMBus adapters for connected hardware monitoring devices. This is the most risky part, and while it works reasonably well on most systems, it has been reported to cause trouble on some systems. Do you want to probe the I2C/SMBus adapters now? (YES/no): y Using driver `i2c-i801' for device 0000:00:1f.3: Intel 3400/5 Series (PCH) Now follows a summary of the probes I have just done. Just press ENTER to continue: Driver `coretemp': * Chip `Intel digital thermal sensor' (confidence: 9) To load everything that is needed, add this to /etc/modules: #----cut here---- # Chip drivers coretemp #----cut here---- If you have some drivers built into your kernel, the list above will contain too many modules. Skip the appropriate ones! Do you want to add these lines automatically to /etc/modules? (yes/NO) My output for 'sensors' is: acpitz-virtual-0 Adapter: Virtual device temp1: +58.0°C (crit = +100.0°C) coretemp-isa-0000 Adapter: ISA adapter Core 0: +56.0°C (high = +84.0°C, crit = +100.0°C) Core 1: +57.0°C (high = +84.0°C, crit = +100.0°C) Core 2: +58.0°C (high = +84.0°C, crit = +100.0°C) Core 3: +57.0°C (high = +84.0°C, crit = +100.0°C) and my '/etc/modules' is: # /etc/modules: kernel modules to load at boot time. # # This file contains the names of kernel modules that should be loaded # at boot time, one per line. Lines beginning with "#" are ignored. lp rtc # Generated by sensors-detect on Fri Nov 30 23:24:31 2012 # Chip drivers coretemp

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• IBM "per core" comparisons for SPECjEnterprise2010

  - by jhenning

  I recently stumbled upon a blog entry from Roman Kharkovski (an IBM employee) comparing some SPECjEnterprise2010 results for IBM vs. Oracle. Mr. Kharkovski's blog claims that SPARC delivers half the transactions per core vs. POWER7. Prior to any argument, I should say that my predisposition is to like Mr. Kharkovski, because he says that his blog is intended to be factual; that the intent is to try to avoid marketing hype and FUD tactic; and mostly because he features a picture of himself wearing a bike helmet (me too). Therefore, in a spirit of technical argument, rather than FUD fight, there are a few areas in his comparison that should be discussed. Scaling is not free For any benchmark, if a small system scores 13k using quantity R1 of some resource, and a big system scores 57k using quantity R2 of that resource, then, sure, it's tempting to divide: is  13k/R1 > 57k/R2 ? It is tempting, but not necessarily educational. The problem is that scaling is not free. Building big systems is harder than building small systems. Scoring  13k/R1  on a little system provides no guarantee whatsoever that one can sustain that ratio when attempting to handle more than 4 times as many users. Choosing the denominator radically changes the picture When ratios are used, one can vastly manipulate appearances by the choice of denominator. In this case, lots of choices are available for the resource to be compared (R1 and R2 above). IBM chooses to put cores in the denominator. Mr. Kharkovski provides some reasons for that choice in his blog entry. And yet, it should be noted that the very concept of a core is: arbitrary: not necessarily comparable across vendors; fluid: modern chips shift chip resources in response to load; and invisible: unless you have a microscope, you can't see it. By contrast, one can actually see processor chips with the naked eye, and they are a bit easier to count. If we put chips in the denominator instead of cores, we get: 13161.07 EjOPS / 4 chips = 3290 EjOPS per chip for IBM vs 57422.17 EjOPS / 16 chips = 3588 EjOPS per chip for Oracle The choice of denominator makes all the difference in the appearance. Speaking for myself, dividing by chips just seems to make more sense, because: I can see chips and count them; and I can accurately compare the number of chips in my system to the count in some other vendor's system; and Tthe probability of being able to continue to accurately count them over the next 10 years of microprocessor development seems higher than the probability of being able to accurately and comparably count "cores". SPEC Fair use requirements Speaking as an individual, not speaking for SPEC and not speaking for my employer, I wonder whether Mr. Kharkovski's blog article, taken as a whole, meets the requirements of the SPEC Fair Use rule www.spec.org/fairuse.html section I.D.2. For example, Mr. Kharkovski's footnote (1) begins Results from http://www.spec.org as of 04/04/2013 Oracle SUN SPARC T5-8 449 EjOPS/core SPECjEnterprise2010 (Oracle's WLS best SPECjEnterprise2010 EjOPS/core result on SPARC). IBM Power730 823 EjOPS/core (World Record SPECjEnterprise2010 EJOPS/core result) The questionable tactic, from a Fair Use point of view, is that there is no such metric at the designated location. At www.spec.org, You can find the SPEC metric 57422.17 SPECjEnterprise2010 EjOPS for Oracle and You can also find the SPEC metric 13161.07 SPECjEnterprise2010 EjOPS for IBM. Despite the implication of the footnote, you will not find any mention of 449 nor anything that says 823. SPEC says that you can, under its fair use rule, derive your own values; but it emphasizes: "The context must not give the appearance that SPEC has created or endorsed the derived value." Substantiation and transparency Although SPEC disclaims responsibility for non-SPEC information (section I.E), it says that non-SPEC data and methods should be accurate, should be explained, should be substantiated. Unfortunately, it is difficult or impossible for the reader to independently verify the pricing: Were like units compared to like (e.g. list price to list price)? Were all components (hw, sw, support) included? Were all fees included? Note that when tpc.org shows IBM pricing, there are often items such as "PROCESSOR ACTIVATION" and "MEMORY ACTIVATION". Without the transparency of a detailed breakdown, the pricing claims are questionable. T5 claim for "Fastest Processor" Mr. Kharkovski several times questions Oracle's claim for fastest processor, writing You see, when you publish industry benchmarks, people may actually compare your results to other vendor's results. Well, as we performance people always say, "it depends". If you believe in performance-per-core as the primary way of looking at the world, then yes, the POWER7+ is impressive, spending its chip resources to support up to 32 threads (8 cores x 4 threads). Or, it just might be useful to consider performance-per-chip. Each SPARC T5 chip allows 128 hardware threads to be simultaneously executing (16 cores x 8 threads). The Industry Standard Benchmark that focuses specifically on processor chip performance is SPEC CPU2006. For this very well known and popular benchmark, SPARC T5: provides better performance than both POWER7 and POWER7+, for 1 chip vs. 1 chip, for 8 chip vs. 8 chip, for integer (SPECint_rate2006) and floating point (SPECfp_rate2006), for Peak tuning and for Base tuning. For example, at the 8-chip level, integer throughput (SPECint_rate2006) is: 3750 for SPARC 2170 for POWER7+. You can find the details at the March 2013 BestPerf CPU2006 page SPEC is a trademark of the Standard Performance Evaluation Corporation, www.spec.org. The two specific results quoted for SPECjEnterprise2010 are posted at the URLs linked from the discussion. Results for SPEC CPU2006 were verified at spec.org 1 July 2013, and can be rechecked here.

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• Cant get lm-sensors to load ATI Radeon temp and fan or output all settings

  - by woody

  New to Linux and having minor issues :/ . I followed this guide initially but did not recieve the proper output and did not show my ATI Radeon HD 5000 temp or fan speed. Then used this guide, same problems exhibited. No issues installing and no errors. I think its not reading i2c for some reason. The proprietary driver is installed and functioning correctly according fglrxinfo. I can use aticonfig commands and view both temp and fan. Any ideas on how to get the ATI Radeon sensors working under 'sensors'? When i run 'sudo sensors-detect' this is my ouput # sensors-detect revision 5984 (2011-07-10 21:22:53 +0200) # System: LENOVO IdeaPad Y560 (laptop) # Board: Lenovo KL3 This program will help you determine which kernel modules you need to load to use lm_sensors most effectively. It is generally safe and recommended to accept the default answers to all questions, unless you know what you're doing. Some south bridges, CPUs or memory controllers contain embedded sensors. Do you want to scan for them? This is totally safe. (YES/no): y Silicon Integrated Systems SIS5595... No VIA VT82C686 Integrated Sensors... No VIA VT8231 Integrated Sensors... No AMD K8 thermal sensors... No AMD Family 10h thermal sensors... No AMD Family 11h thermal sensors... No AMD Family 12h and 14h thermal sensors... No AMD Family 15h thermal sensors... No AMD Family 15h power sensors... No Intel digital thermal sensor... Success! (driver `coretemp') Intel AMB FB-DIMM thermal sensor... No VIA C7 thermal sensor... No VIA Nano thermal sensor... No Some Super I/O chips contain embedded sensors. We have to write to standard I/O ports to probe them. This is usually safe. Do you want to scan for Super I/O sensors? (YES/no): y Probing for Super-I/O at 0x2e/0x2f Trying family `National Semiconductor/ITE'... Yes Found unknown chip with ID 0x8502 Probing for Super-I/O at 0x4e/0x4f Trying family `National Semiconductor/ITE'... No Trying family `SMSC'... No Trying family `VIA/Winbond/Nuvoton/Fintek'... No Trying family `ITE'... No Some hardware monitoring chips are accessible through the ISA I/O ports. We have to write to arbitrary I/O ports to probe them. This is usually safe though. Yes, you do have ISA I/O ports even if you do not have any ISA slots! Do you want to scan the ISA I/O ports? (YES/no): y Probing for `National Semiconductor LM78' at 0x290... No Probing for `National Semiconductor LM79' at 0x290... No Probing for `Winbond W83781D' at 0x290... No Probing for `Winbond W83782D' at 0x290... No Lastly, we can probe the I2C/SMBus adapters for connected hardware monitoring devices. This is the most risky part, and while it works reasonably well on most systems, it has been reported to cause trouble on some systems. Do you want to probe the I2C/SMBus adapters now? (YES/no): y Using driver `i2c-i801' for device 0000:00:1f.3: Intel 3400/5 Series (PCH) Now follows a summary of the probes I have just done. Just press ENTER to continue: Driver `coretemp': * Chip `Intel digital thermal sensor' (confidence: 9) To load everything that is needed, add this to /etc/modules: #----cut here---- # Chip drivers coretemp #----cut here---- If you have some drivers built into your kernel, the list above will contain too many modules. Skip the appropriate ones! Do you want to add these lines automatically to /etc/modules? (yes/NO) My output for 'sensors' is: acpitz-virtual-0 Adapter: Virtual device temp1: +58.0°C (crit = +100.0°C) coretemp-isa-0000 Adapter: ISA adapter Core 0: +56.0°C (high = +84.0°C, crit = +100.0°C) Core 1: +57.0°C (high = +84.0°C, crit = +100.0°C) Core 2: +58.0°C (high = +84.0°C, crit = +100.0°C) Core 3: +57.0°C (high = +84.0°C, crit = +100.0°C) and my '/etc/modules' is: # /etc/modules: kernel modules to load at boot time. # # This file contains the names of kernel modules that should be loaded # at boot time, one per line. Lines beginning with "#" are ignored. lp rtc # Generated by sensors-detect on Fri Nov 30 23:24:31 2012 # Chip drivers coretemp

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• Writing an ASP.Net Web based TFS Client

  - by Glav

  So one of the things I needed to do was write an ASP.Net MVC based application for our senior execs to manage a set of arbitrary attributes against stories, bugs etc to be able to attribute whether the item was related to Research and Development, and if so, what kind. We are using TFS Azure and don’t have the option of custom templates. I have decided on using a string based field within the template that is not very visible and which we don’t use to write a small set of custom which will determine the research and development association. However, this string munging on the field is not very user friendly so we need a simple tool that can display attributes against items in a simple dropdown list or something similar. Enter a custom web app that accesses our TFS items in Azure (Note: We are also using Visual Studio 2012) Now TFS Azure uses your Live ID and it is not really possible to easily do this in a server based app where no interaction is available. Even if you capture the Live ID credentials yourself and try to submit them to TFS Azure, it wont work. Bottom line is that it is not straightforward nor obvious what you have to do. In fact, it is a real pain to find and there are some answers out there which don’t appear to be answers at all given they didn’t work in my scenario. So for anyone else who wants to do this, here is a simple breakdown on what you have to do: Go here and get the “TFS Service Credential Viewer”. Install it, run it and connect to your TFS instance in azure and create a service account. Note the username and password exactly as it presents it to you. This is the magic identity that will allow unattended, programmatic access. Without this step, don’t bother trying to do anything else. In your MVC app, reference the following assemblies from “C:\Program Files (x86)\Microsoft Visual Studio 11.0\Common7\IDE\ReferenceAssemblies\v2.0”: Microsoft.TeamFoundation.Client.dll Microsoft.TeamFoundation.Common.dll Microsoft.TeamFoundation.VersionControl.Client.dll Microsoft.TeamFoundation.VersionControl.Common.dll Microsoft.TeamFoundation.WorkItemTracking.Client.DataStoreLoader.dll Microsoft.TeamFoundation.WorkItemTracking.Client.dll Microsoft.TeamFoundation.WorkItemTracking.Common.dll If hosting this in Internet Information Server, for the application pool this app runs under, you will need to enable 32 Bit support. You also have to allow the TFS client assemblies to store a cache of files on your system. If you don’t do this, you will authenticate fine, but then get an exception saying that it is unable to access the cache at some directory path when you query work items. You can set this up by adding the following to your web.config, in the <appSettings> element as shown below: <appSettings> <!-- Add reference to TFS Client Cache --> <add key="WorkItemTrackingCacheRoot" value="C:\windows\temp" /> </appSettings> With all that in place, you can write the following code: var token = new Microsoft.TeamFoundation.Client.SimpleWebTokenCredential("{you-service-account-name", "{your-service-acct-password}"); var clientCreds = new Microsoft.TeamFoundation.Client.TfsClientCredentials(token); var currentCollection = new TfsTeamProjectCollection(new Uri(“https://{yourdomain}.visualstudio.com/defaultcollection”), clientCreds); TfsConfigurationServercurrentCollection.EnsureAuthenticated(); In the above code, not the URL contains the “defaultcollection” at the end of the URL. Obviously replace {yourdomain} with whatever is defined for your TFS in Azure instance. In addition, make sure the service user account and password that was generated in the first step is substituted in here. Note: If something is not right, the “EnsureAuthenticated()” call will throw an exception with the message being you are not authorised. If you forget the “defaultcollection” on the URL, it will still fail but with a message saying you are not authorised. That is, a similar but different exception message. And that is it. You can then query the collection using something like: var service = currentCollection.GetService<WorkItemStore>(); var proj = service.Projects[0]; var allQueries = proj.StoredQueries; for (int qcnt = 0; qcnt < allQueries.Count; qcnt++) {     var query = allQueries[qcnt];     var queryDesc = string.format(“Query found named: {0}”,query.Name); } You get the idea. If you search around, you will find references to the ServiceIdentityCredentialProvider which is referenced in this article. I had no luck with this method and it all looked too hard since it required an extra KB article and other magic sauce. So I hope that helps. This article certainly would have helped me save a boat load of time and frustration.

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• ROracle support for TimesTen In-Memory Database

  - by Sam Drake

  Today's guest post comes from Jason Feldhaus, a Consulting Member of Technical Staff in the TimesTen Database organization at Oracle.  He shares with us a sample session using ROracle with the TimesTen In-Memory database.  Beginning in version 1.1-4, ROracle includes support for the Oracle Times Ten In-Memory Database, version 11.2.2. TimesTen is a relational database providing very fast and high throughput through its memory-centric architecture.  TimesTen is designed for low latency, high-volume data, and event and transaction management. A TimesTen database resides entirely in memory, so no disk I/O is required for transactions and query operations. TimesTen is used in applications requiring very fast and predictable response time, such as real-time financial services trading applications and large web applications. TimesTen can be used as the database of record or as a relational cache database to Oracle Database. ROracle provides an interface between R and the database, providing the rich functionality of the R statistical programming environment using the SQL query language. ROracle uses the OCI libraries to handle database connections, providing much better performance than standard ODBC.The latest ROracle enhancements include: Support for Oracle TimesTen In-Memory Database Support for Date-Time using R's POSIXct/POSIXlt data types RAW, BLOB and BFILE data type support Option to specify number of rows per fetch operation Option to prefetch LOB data Break support using Ctrl-C Statement caching support Times Ten 11.2.2 contains enhanced support for analytics workloads and complex queries: Analytic functions: AVG, SUM, COUNT, MAX, MIN, DENSE_RANK, RANK, ROW_NUMBER, FIRST_VALUE and LAST_VALUE Analytic clauses: OVER PARTITION BY and OVER ORDER BY Multidimensional grouping operators: Grouping clauses: GROUP BY CUBE, GROUP BY ROLLUP, GROUP BY GROUPING SETS Grouping functions: GROUP, GROUPING_ID, GROUP_ID WITH clause, which allows repeated references to a named subquery block Aggregate expressions over DISTINCT expressions General expressions that return a character string in the source or a pattern within the LIKE predicate Ability to order nulls first or last in a sort result (NULLS FIRST or NULLS LAST in the ORDER BY clause) Note: Some functionality is only available with Oracle Exalytics, refer to the TimesTen product licensing document for details. Connecting to TimesTen is easy with ROracle. Simply install and load the ROracle package and load the driver. > install.packages("ROracle") > library(ROracle) Loading required package: DBI > drv <- dbDriver("Oracle") Once the ROracle package is installed, create a database connection object and connect to a TimesTen direct driver DSN as the OS user. > conn <- dbConnect(drv, username ="", password="", dbname = "localhost/SampleDb_1122:timesten_direct") You have the option to report the server type - Oracle or TimesTen? > print (paste ("Server type =", dbGetInfo (conn)$serverType)) [1] "Server type = TimesTen IMDB" To create tables in the database using R data frame objects, use the function dbWriteTable. In the following example we write the built-in iris data frame to TimesTen. The iris data set is a small example data set containing 150 rows and 5 columns. We include it here not to highlight performance, but so users can easily run this example in their R session. > dbWriteTable (conn, "IRIS", iris, overwrite=TRUE, ora.number=FALSE) [1] TRUE Verify that the newly created IRIS table is available in the database. To list the available tables and table columns in the database, use dbListTables and dbListFields, respectively. > dbListTables (conn) [1] "IRIS" > dbListFields (conn, "IRIS") [1] "SEPAL.LENGTH" "SEPAL.WIDTH" "PETAL.LENGTH" "PETAL.WIDTH" "SPECIES" To retrieve a summary of the data from the database we need to save the results to a local object. The following call saves the results of the query as a local R object, iris.summary. The ROracle function dbGetQuery is used to execute an arbitrary SQL statement against the database. When connected to TimesTen, the SQL statement is processed completely within main memory for the fastest response time. > iris.summary <- dbGetQuery(conn, 'SELECT SPECIES, AVG ("SEPAL.LENGTH") AS AVG_SLENGTH, AVG ("SEPAL.WIDTH") AS AVG_SWIDTH, AVG ("PETAL.LENGTH") AS AVG_PLENGTH, AVG ("PETAL.WIDTH") AS AVG_PWIDTH FROM IRIS GROUP BY ROLLUP (SPECIES)') > iris.summary SPECIES AVG_SLENGTH AVG_SWIDTH AVG_PLENGTH AVG_PWIDTH 1 setosa 5.006000 3.428000 1.462 0.246000 2 versicolor 5.936000 2.770000 4.260 1.326000 3 virginica 6.588000 2.974000 5.552 2.026000 4 <NA> 5.843333 3.057333 3.758 1.199333 Finally, disconnect from the TimesTen Database. > dbCommit (conn) [1] TRUE > dbDisconnect (conn) [1] TRUE We encourage you download Oracle software for evaluation from the Oracle Technology Network. See these links for our software: Times Ten In-Memory Database,  ROracle.  As always, we welcome comments and questions on the TimesTen and  Oracle R technical forums.

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• SQL ADO.NET shortcut extensions (old school!)

  - by Jeff

  As much as I love me some ORM's (I've used LINQ to SQL quite a bit, and for the MSDN/TechNet Profile and Forums we're using NHibernate more and more), there are times when it's appropriate, and in some ways more simple, to just throw up so old school ADO.NET connections, commands, readers and such. It still feels like a pain though to new up all the stuff, make sure it's closed, blah blah blah. It's pretty much the least favorite task of writing data access code. To minimize the pain, I have a set of extension methods that I like to use that drastically reduce the code you have to write. Here they are... public static void Using(this SqlConnection connection, Action<SqlConnection> action) {     connection.Open();     action(connection);     connection.Close(); } public static SqlCommand Command(this SqlConnection connection, string sql){    var command = new SqlCommand(sql, connection);    return command;}public static SqlCommand AddParameter(this SqlCommand command, string parameterName, object value){    command.Parameters.AddWithValue(parameterName, value);    return command;}public static object ExecuteAndReturnIdentity(this SqlCommand command){    if (command.Connection == null)        throw new Exception("SqlCommand has no connection.");    command.ExecuteNonQuery();    command.Parameters.Clear();    command.CommandText = "SELECT @@IDENTITY";    var result = command.ExecuteScalar();    return result;}public static SqlDataReader ReadOne(this SqlDataReader reader, Action<SqlDataReader> action){    if (reader.Read())        action(reader);    reader.Close();    return reader;}public static SqlDataReader ReadAll(this SqlDataReader reader, Action<SqlDataReader> action){    while (reader.Read())        action(reader);    reader.Close();    return reader;} It has been awhile since I've really revisited these, so you will likely find opportunity for further optimization. The bottom line here is that you can chain together a bunch of these methods to make a much more concise database call, in terms of the code on your screen, anyway. Here are some examples: public Dictionary<string, string> Get(){    var dictionary = new Dictionary<string, string>();    _sqlHelper.GetConnection().Using(connection =>        connection.Command("SELECT Setting, [Value] FROM Settings")            .ExecuteReader()            .ReadAll(r => dictionary.Add(r.GetString(0), r.GetString(1))));    return dictionary;} or... public void ChangeName(User user, string newName){    _sqlHelper.GetConnection().Using(connection =>         connection.Command("UPDATE Users SET Name = @Name WHERE UserID = @UserID")            .AddParameter("@Name", newName)            .AddParameter("@UserID", user.UserID)            .ExecuteNonQuery());} The _sqlHelper.GetConnection() is just some other code that gets a connection object for you. You might have an even cleaner way to take that step out entirely. This looks more fluent, and the real magic sauce for me is the reader bits where you can put any kind of arbitrary method in there to iterate over the results.

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• Windows in StreamInsight: Hopping vs. Snapshot

  - by Roman Schindlauer

  Three weeks ago, we explained the basic concept of windows in StreamInsight: defining sets of events that serve as arguments for set-based operations, like aggregations. Today, we want to discuss the so-called Hopping Windows and compare them with Snapshot Windows. We will compare these two, because they can serve similar purposes with different behaviors; we will discuss the remaining window type, Count Windows, another time. Hopping (and its syntactic-sugar-sister Tumbling) windows are probably the most straightforward windowing concept in StreamInsight. A hopping window is defined by its length, and the offset from one window to the next. They are aligned with some absolute point on the timeline (which can also be given as a parameter to the window) and create sets of events. The diagram below shows an example of a hopping window with length of 1h and hop size (the offset) of 15 minutes, hence creating overlapping windows:   Two aspects in this diagram are important: Since this window is overlapping, an event can fall into more than one windows. If an (interval) event spans a window boundary, its lifetime will be clipped to the window, before it is passed to the set-based operation. That’s the default and currently only available window input policy. (This should only concern you if you are using a time-sensitive user-defined aggregate or operator.) The set-based operation will be applied to each of these sets, yielding a result. This result is: A single scalar value in case of built-in or user-defined aggregates. A subset of the input payloads, in case of the TopK operator. Arbitrary events, when using a user-defined operator. The timestamps of the result are almost always the ones of the windows. Only the user-defined  operator can create new events with timestamps. (However, even these event lifetimes are subject to the window’s output policy, which is currently always to clip to the window end.) Let’s assume we were calculating the sum over some payload field: var result = from window in source.HoppingWindow( TimeSpan.FromHours(1), TimeSpan.FromMinutes(15), HoppingWindowOutputPolicy.ClipToWindowEnd) select new { avg = window.Avg(e => e.Value) }; Now each window is reflected by one result event:   As you can see, the window definition defines the output frequency. No matter how many or few events we got from the input, this hopping window will produce one result every 15 minutes – except for those windows that do not contain any events at all, because StreamInsight window operations are empty-preserving (more about that another time). The “forced” output for every window can become a performance issue if you have a real-time query with many events in a wide group & apply – let me explain: imagine you have a lot of events that you group by and then aggregate within each group – classical streaming pattern. The hopping window produces a result in each group at exactly the same point in time for all groups, since the window boundaries are aligned with the timeline, not with the event timestamps. This means that the query output will become very bursty, delivering the results of all the groups at the same point in time. This becomes especially obvious if the events are long-lasting, spanning multiple windows each, so that the produced result events do not change their value very often. In such a case, a snapshot window can remedy. Snapshot windows are more difficult to explain than hopping windows: they represent those periods in time, when no event changes occur. In other words, if you mark all event start and and times on your timeline, then you are looking at all snapshot window boundaries:   If your events are never overlapping, the snapshot window will not make much sense. It is commonly used together with timestamp modification, which make it a very powerful tool. Or as Allan Mitchell expressed in in a recent tweet: “I used to look at SnapshotWindow() with disdain. Now she is my mistress, the one I turn to in times of trouble and need”. Let’s look at a simple example: I want to compute the average of some value in my events over the last minute. I don’t want this output be produced at fixed intervals, but at soon as it changes (that’s the true event-driven spirit!). The snapshot window will include all currently active event at each point in time, hence we need to extend our original events’ lifetimes into the future: Applying the Snapshot window on these events, it will appear to be “looking back into the past”: If you look at the result produced in this diagram, you can easily prove that, at each point in time, the current event value represents the average of all original input event within the last minute. Here is the LINQ representation of that query, applying the lifetime extension before the snapshot window: var result = from window in source .AlterEventDuration(e => TimeSpan.FromMinutes(1)) .SnapshotWindow(SnapshotWindowOutputPolicy.Clip) select new { avg = window.Avg(e => e.Value) }; With more complex modifications of the event lifetimes you can achieve many more query patterns. For instance “running totals” by keeping the event start times, but snapping their end times to some fixed time, like the end of the day. Each snapshot then “sees” all events that have happened in the respective time period so far. Regards, The StreamInsight Team

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• Career guidance/advice for Junior-level Software Engineer [closed]

  - by John Do

  I have quite a few questions on my mind, so please bare with me. Please don't feel obligated to answer all of them, any as you choose will do. I'd appreciate if you could share some insight on any of these. Before I begin, some context: I currently have almost two years of professional experience as a Software Engineer, mainly developing software in Java. At this point, I feel that I have reached the peak in my career growth at the current company I am at and therefore I am looking for a new job, ideally again, as a Software Engineer. I have been interviewing for the past few months casually but have not had luck with companies I have a passion for. So, in no particular order - 1) In general, what are your thoughts on having graduate degrees in CS / Software Engineering. How much does it influence a salary increase, and do you think it's beneficial when working on real-world problems? I get the sense that a graduate degree in the field is trivial unless you really have a passion for research. 2) In general, in professional practice, how often had you have to write your own data structures and "complex" algorithms from scratch? In my own work, I have found myself relying mainly on third-party frameworks and the Java standard library to implement solutions as per business requirements. What are your thoughts on this? 3) In terms of resume, I feel the most ambivalent here. I want to be able to "blemish" my resume to a certain extent so that it stands out from others', but at the same time I do not want to over-exagerate my abilities. How do you strike a balance here? For example: I say that I am proficient in Java with data structures and algorithms. This is obviously a subjective and relative statement. I've taken the classes in my undergrad, and I've applied it in my work experience. What I feel as "prociency" can be seen as junior-level to others. How do you know what to say? Most of the time, recruiters (with no technical background) will be looking for keywords that stand out. This leads me to my next question (4). 4) Just from interviewing for the past few months (and getting plenty of rejections), I've come to realize that I may not be as proficient in data structures and algorithms as I thought I was. Do you think it's a good idea to remove the "proficient in java/data structure and algorithms"? I feel that being too hoenst on the resume will impede me from scoring opportunities to even have an interview with top-notch companies. What are your thoughts? 5) What is the absolute "must-have" knowledge going into a technical interview? I've been practicing several algorithmic and data sturcture problems now, and I feel that my abilities to solve arbitrary problems efficiently has not gotten significantly better. Do you think these abilities are something innate - it's either you have in you, or you don't? How can you teach yourself to learn, if you will? 6) How easy is it to go from industry/function to the next? I work mainly with backend technologies and I'm now interested in working with the frontend, i.e javascript,jquery,php or even mobile development. In your own experience, how did you not get pidgeon holed in your career? I feel that the choices you make now ultimately decide your future. As cliche as it sounds, I think it may be true. Here's what I mean: you've worked mainly as a backend engineer, people are interested in you doing the same thing since you've already accumulated experience in that function. How do get experience in a new function if people won't accept you because you don't already have it? It's a catch 22, you see... Are side projects the only real way to help you move from one function to another that you're truly interested in? For example: I could start writing my own mobile applications, even though I've worked mainly on the backend. Thanks so much for the long read. As a relatively new engineer to the real world, I am very humble and would like those who are experienced to shed some light. Thank you so much.

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

  - by Compudicted

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

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• ROracle support for TimesTen In-Memory Database

  - by Sherry LaMonica

  Today's guest post comes from Jason Feldhaus, a Consulting Member of Technical Staff in the TimesTen Database organization at Oracle.  He shares with us a sample session using ROracle with the TimesTen In-Memory database.  Beginning in version 1.1-4, ROracle includes support for the Oracle Times Ten In-Memory Database, version 11.2.2. TimesTen is a relational database providing very fast and high throughput through its memory-centric architecture.  TimesTen is designed for low latency, high-volume data, and event and transaction management. A TimesTen database resides entirely in memory, so no disk I/O is required for transactions and query operations. TimesTen is used in applications requiring very fast and predictable response time, such as real-time financial services trading applications and large web applications. TimesTen can be used as the database of record or as a relational cache database to Oracle Database. ROracle provides an interface between R and the database, providing the rich functionality of the R statistical programming environment using the SQL query language. ROracle uses the OCI libraries to handle database connections, providing much better performance than standard ODBC.The latest ROracle enhancements include: Support for Oracle TimesTen In-Memory Database Support for Date-Time using R's POSIXct/POSIXlt data types RAW, BLOB and BFILE data type support Option to specify number of rows per fetch operation Option to prefetch LOB data Break support using Ctrl-C Statement caching support Times Ten 11.2.2 contains enhanced support for analytics workloads and complex queries: Analytic functions: AVG, SUM, COUNT, MAX, MIN, DENSE_RANK, RANK, ROW_NUMBER, FIRST_VALUE and LAST_VALUE Analytic clauses: OVER PARTITION BY and OVER ORDER BY Multidimensional grouping operators: Grouping clauses: GROUP BY CUBE, GROUP BY ROLLUP, GROUP BY GROUPING SETS Grouping functions: GROUP, GROUPING_ID, GROUP_ID WITH clause, which allows repeated references to a named subquery block Aggregate expressions over DISTINCT expressions General expressions that return a character string in the source or a pattern within the LIKE predicate Ability to order nulls first or last in a sort result (NULLS FIRST or NULLS LAST in the ORDER BY clause) Note: Some functionality is only available with Oracle Exalytics, refer to the TimesTen product licensing document for details. Connecting to TimesTen is easy with ROracle. Simply install and load the ROracle package and load the driver. > install.packages("ROracle") > library(ROracle) Loading required package: DBI > drv <- dbDriver("Oracle") Once the ROracle package is installed, create a database connection object and connect to a TimesTen direct driver DSN as the OS user. > conn <- dbConnect(drv, username ="", password="", dbname = "localhost/SampleDb_1122:timesten_direct") You have the option to report the server type - Oracle or TimesTen? > print (paste ("Server type =", dbGetInfo (conn)$serverType)) [1] "Server type = TimesTen IMDB" To create tables in the database using R data frame objects, use the function dbWriteTable. In the following example we write the built-in iris data frame to TimesTen. The iris data set is a small example data set containing 150 rows and 5 columns. We include it here not to highlight performance, but so users can easily run this example in their R session. > dbWriteTable (conn, "IRIS", iris, overwrite=TRUE, ora.number=FALSE) [1] TRUE Verify that the newly created IRIS table is available in the database. To list the available tables and table columns in the database, use dbListTables and dbListFields, respectively. > dbListTables (conn) [1] "IRIS" > dbListFields (conn, "IRIS") [1] "SEPAL.LENGTH" "SEPAL.WIDTH" "PETAL.LENGTH" "PETAL.WIDTH" "SPECIES" To retrieve a summary of the data from the database we need to save the results to a local object. The following call saves the results of the query as a local R object, iris.summary. The ROracle function dbGetQuery is used to execute an arbitrary SQL statement against the database. When connected to TimesTen, the SQL statement is processed completely within main memory for the fastest response time. > iris.summary <- dbGetQuery(conn, 'SELECT SPECIES, AVG ("SEPAL.LENGTH") AS AVG_SLENGTH, AVG ("SEPAL.WIDTH") AS AVG_SWIDTH, AVG ("PETAL.LENGTH") AS AVG_PLENGTH, AVG ("PETAL.WIDTH") AS AVG_PWIDTH FROM IRIS GROUP BY ROLLUP (SPECIES)') > iris.summary SPECIES AVG_SLENGTH AVG_SWIDTH AVG_PLENGTH AVG_PWIDTH 1 setosa 5.006000 3.428000 1.462 0.246000 2 versicolor 5.936000 2.770000 4.260 1.326000 3 virginica 6.588000 2.974000 5.552 2.026000 4 <NA> 5.843333 3.057333 3.758 1.199333 Finally, disconnect from the TimesTen Database. > dbCommit (conn) [1] TRUE > dbDisconnect (conn) [1] TRUE We encourage you download Oracle software for evaluation from the Oracle Technology Network. See these links for our software: Times Ten In-Memory Database,  ROracle.  As always, we welcome comments and questions on the TimesTen and  Oracle R technical forums.

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• Breaking 1NF to model subset constraints. Does this sound sane?

  - by Chris Travers

  My first question here. Appologize if it is in the wrong forum but this seems pretty conceptual. I am looking at doing something that goes against conventional wisdom and want to get some feedback as to whether this is totally insane or will result in problems, so critique away! I am on PostgreSQL 9.1 but may be moving to 9.2 for this part of this project. To re-iterate: Does it seem sane to break 1NF in this way? I am not looking for debugging code so much as where people see problems that this might lead. The Problem In double entry accounting, financial transactions are journal entries with an arbitrary number of lines. Each line has either a left value (debit) or a right value (credit) which can be modelled as a single value with negatives as debits and positives as credits or vice versa. The sum of all debits and credits must equal zero (so if we go with a single amount field, sum(amount) must equal zero for each financial journal entry). SQL-based databases, pretty much required for this sort of work, have no way to express this sort of constraint natively and so any approach to enforcing it in the database seems rather complex. The Write Model The journal entries are append only. There is a possibility we will add a delete model but it will be subject to a different set of restrictions and so is not applicable here. If and when we allow deletes, we will probably do them using a simple ON DELETE CASCADE designation on the foreign key, and require that deletes go through a dedicated stored procedure which can enforce the other constraints. So inserts and selects have to be accommodated but updates and deletes do not for this task. My Proposed Solution My proposed solution is to break first normal form and model constraints on arrays of tuples, with a trigger that breaks the rows out into another table. CREATE TABLE journal_line ( entry_id bigserial primary key, account_id int not null references account(id), journal_entry_id bigint not null, -- adding references later amount numeric not null ); I would then add "table methods" to extract debits and credits for reporting purposes: CREATE OR REPLACE FUNCTION debits(journal_line) RETURNS numeric LANGUAGE sql IMMUTABLE AS $$ SELECT CASE WHEN $1.amount < 0 THEN $1.amount * -1 ELSE NULL END; $$; CREATE OR REPLACE FUNCTION credits(journal_line) RETURNS numeric LANGUAGE sql IMMUTABLE AS $$ SELECT CASE WHEN $1.amount > 0 THEN $1.amount ELSE NULL END; $$; Then the journal entry table (simplified for this example): CREATE TABLE journal_entry ( entry_id bigserial primary key, -- no natural keys :-( journal_id int not null references journal(id), date_posted date not null, reference text not null, description text not null, journal_lines journal_line[] not null ); Then a table method and and check constraints: CREATE OR REPLACE FUNCTION running_total(journal_entry) returns numeric language sql immutable as $$ SELECT sum(amount) FROM unnest($1.journal_lines); $$; ALTER TABLE journal_entry ADD CONSTRAINT CHECK (((journal_entry.running_total) = 0)); ALTER TABLE journal_line ADD FOREIGN KEY journal_entry_id REFERENCES journal_entry(entry_id); And finally we'd have a breakout trigger: CREATE OR REPLACE FUNCTION je_breakout() RETURNS TRIGGER LANGUAGE PLPGSQL AS $$ BEGIN IF TG_OP = 'INSERT' THEN INSERT INTO journal_line (journal_entry_id, account_id, amount) SELECT NEW.id, account_id, amount FROM unnest(NEW.journal_lines); RETURN NEW; ELSE RAISE EXCEPTION 'Operation Not Allowed'; END IF; END; $$; And finally CREATE TRIGGER AFTER INSERT OR UPDATE OR DELETE ON journal_entry FOR EACH ROW EXECUTE_PROCEDURE je_breaout(); Of course the example above is simplified. There will be a status table that will track approval status allowing for separation of duties, etc. However the goal here is to prevent unbalanced transactions. Any feedback? Does this sound entirely insane? Standard Solutions? In getting to this point I have to say I have looked at four different current ERP solutions to this problems: Represent every line item as a debit and a credit against different accounts. Use of foreign keys against the line item table to enforce an eventual running total of 0 Use of constraint triggers in PostgreSQL Forcing all validation here solely through the app logic. My concerns are that #1 is pretty limiting and very hard to audit internally. It's not programmer transparent and so it strikes me as being difficult to work with in the future. The second strikes me as being very complex and required a series of contraints and foreign keys against self to make work, and therefore it strikes me as complex, hard to sort out at least in my mind, and thus hard to work with. The fourth could be done as we force all access through stored procedures anyway and this is the most common solution (have the app total things up and throw an error otherwise). However, I think proof that a constraint is followed is superior to test cases, and so the question becomes whether this in fact generates insert anomilies rather than solving them. If this is a solved problem it isn't the case that everyone agrees on the solution....

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• How can I estimate the entropy of a password?

  - by Wug

  Having read various resources about password strength I'm trying to create an algorithm that will provide a rough estimation of how much entropy a password has. I'm trying to create an algorithm that's as comprehensive as possible. At this point I only have pseudocode, but the algorithm covers the following: password length repeated characters patterns (logical) different character spaces (LC, UC, Numeric, Special, Extended) dictionary attacks It does NOT cover the following, and SHOULD cover it WELL (though not perfectly): ordering (passwords can be strictly ordered by output of this algorithm) patterns (spatial) Can anyone provide some insight on what this algorithm might be weak to? Specifically, can anyone think of situations where feeding a password to the algorithm would OVERESTIMATE its strength? Underestimations are less of an issue. The algorithm: // the password to test password = ? length = length(password) // unique character counts from password (duplicates discarded) uqlca = number of unique lowercase alphabetic characters in password uquca = number of uppercase alphabetic characters uqd = number of unique digits uqsp = number of unique special characters (anything with a key on the keyboard) uqxc = number of unique special special characters (alt codes, extended-ascii stuff) // algorithm parameters, total sizes of alphabet spaces Nlca = total possible number of lowercase letters (26) Nuca = total uppercase letters (26) Nd = total digits (10) Nsp = total special characters (32 or something) Nxc = total extended ascii characters that dont fit into other categorys (idk, 50?) // algorithm parameters, pw strength growth rates as percentages (per character) flca = entropy growth factor for lowercase letters (.25 is probably a good value) fuca = EGF for uppercase letters (.4 is probably good) fd = EGF for digits (.4 is probably good) fsp = EGF for special chars (.5 is probably good) fxc = EGF for extended ascii chars (.75 is probably good) // repetition factors. few unique letters == low factor, many unique == high rflca = (1 - (1 - flca) ^ uqlca) rfuca = (1 - (1 - fuca) ^ uquca) rfd = (1 - (1 - fd ) ^ uqd ) rfsp = (1 - (1 - fsp ) ^ uqsp ) rfxc = (1 - (1 - fxc ) ^ uqxc ) // digit strengths strength = ( rflca * Nlca + rfuca * Nuca + rfd * Nd + rfsp * Nsp + rfxc * Nxc ) ^ length entropybits = log_base_2(strength) A few inputs and their desired and actual entropy_bits outputs: INPUT DESIRED ACTUAL aaa very pathetic 8.1 aaaaaaaaa pathetic 24.7 abcdefghi weak 31.2 H0ley$Mol3y_ strong 72.2 s^fU¬5ü;y34G< wtf 88.9 [a^36]* pathetic 97.2 [a^20]A[a^15]* strong 146.8 xkcd1** medium 79.3 xkcd2** wtf 160.5 * these 2 passwords use shortened notation, where [a^N] expands to N a's. ** xkcd1 = "Tr0ub4dor&3", xkcd2 = "correct horse battery staple" The algorithm does realize (correctly) that increasing the alphabet size (even by one digit) vastly strengthens long passwords, as shown by the difference in entropy_bits for the 6th and 7th passwords, which both consist of 36 a's, but the second's 21st a is capitalized. However, they do not account for the fact that having a password of 36 a's is not a good idea, it's easily broken with a weak password cracker (and anyone who watches you type it will see it) and the algorithm doesn't reflect that. It does, however, reflect the fact that xkcd1 is a weak password compared to xkcd2, despite having greater complexity density (is this even a thing?). How can I improve this algorithm? Addendum 1 Dictionary attacks and pattern based attacks seem to be the big thing, so I'll take a stab at addressing those. I could perform a comprehensive search through the password for words from a word list and replace words with tokens unique to the words they represent. Word-tokens would then be treated as characters and have their own weight system, and would add their own weights to the password. I'd need a few new algorithm parameters (I'll call them lw, Nw ~= 2^11, fw ~= .5, and rfw) and I'd factor the weight into the password as I would any of the other weights. This word search could be specially modified to match both lowercase and uppercase letters as well as common character substitutions, like that of E with 3. If I didn't add extra weight to such matched words, the algorithm would underestimate their strength by a bit or two per word, which is OK. Otherwise, a general rule would be, for each non-perfect character match, give the word a bonus bit. I could then perform simple pattern checks, such as searches for runs of repeated characters and derivative tests (take the difference between each character), which would identify patterns such as 'aaaaa' and '12345', and replace each detected pattern with a pattern token, unique to the pattern and length. The algorithmic parameters (specifically, entropy per pattern) could be generated on the fly based on the pattern. At this point, I'd take the length of the password. Each word token and pattern token would count as one character; each token would replace the characters they symbolically represented. I made up some sort of pattern notation, but it includes the pattern length l, the pattern order o, and the base element b. This information could be used to compute some arbitrary weight for each pattern. I'd do something better in actual code. Modified Example: Password: 1234kitty$$$$$herpderp Tokenized: 1 2 3 4 k i t t y $ $ $ $ $ h e r p d e r p Words Filtered: 1 2 3 4 @W5783 $ $ $ $ $ @W9001 @W9002 Patterns Filtered: @P[l=4,o=1,b='1'] @W5783 @P[l=5,o=0,b='$'] @W9001 @W9002 Breakdown: 3 small, unique words and 2 patterns Entropy: about 45 bits, as per modified algorithm Password: correcthorsebatterystaple Tokenized: c o r r e c t h o r s e b a t t e r y s t a p l e Words Filtered: @W6783 @W7923 @W1535 @W2285 Breakdown: 4 small, unique words and no patterns Entropy: 43 bits, as per modified algorithm The exact semantics of how entropy is calculated from patterns is up for discussion. I was thinking something like: entropy(b) * l * (o + 1) // o will be either zero or one The modified algorithm would find flaws with and reduce the strength of each password in the original table, with the exception of s^fU¬5ü;y34G<, which contains no words or patterns.

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• In 3D camera math, calculate what Z depth is pixel unity for a given FOV

  - by badweasel

  I am working in iOS and OpenGL ES 2.0. Through trial and error I've figured out a frustum to where at a specific z depth pixels drawn are 1 to 1 with my source textures. So 1 pixel in my texture is 1 pixel on the screen. For 2d games this is good. Of course it means that I also factor in things like the size of the quad and the size of the texture. For example if my sprite is a quad 32x32 pixels. The quad size is 3.2 units wide and tall. And the texcoords are 32 / the size of the texture wide and tall. Then the frustum is: matrixFrustum(-(float)backingWidth/frustumScale,(float)backingWidth/frustumScale, -(float)backingHeight/frustumScale, (float)backingHeight/frustumScale, 40, 1000, mProjection); Where frustumScale is 800 for a retina screen. Then at a distance of 800 from camera the sprite is pixel for pixel the same as photoshop. For 3d games sometimes I still want to be able to do this. But depending on the scene I sometimes need the FOV to be different things. I'm looking for a way to figure out what Z depth will achieve this same pixel unity for a given FOV. For this my mProjection is set using: matrixPerspective(cameraFOV, near, far, (float)backingWidth / (float)backingHeight, mProjection); With testing I found that at an FOV of 45.0 a Z of 38.5 is very close to pixel unity. And at an FOV of 30.0 a Z of 59.5 is about right. But how can I calculate a value that is spot on? Here's my matrixPerspecitve code: void matrixPerspective(float angle, float near, float far, float aspect, mat4 m) { //float size = near * tanf(angle / 360.0 * M_PI); float size = near * tanf(degreesToRadians(angle) / 2.0); float left = -size, right = size, bottom = -size / aspect, top = size / aspect; // Unused values in perspective formula. m[1] = m[2] = m[3] = m[4] = 0; m[6] = m[7] = m[12] = m[13] = m[15] = 0; // Perspective formula. m[0] = 2 * near / (right - left); m[5] = 2 * near / (top - bottom); m[8] = (right + left) / (right - left); m[9] = (top + bottom) / (top - bottom); m[10] = -(far + near) / (far - near); m[11] = -1; m[14] = -(2 * far * near) / (far - near); } And my mView is set using: lookAtMatrix(cameraPos, camLookAt, camUpVector, mView); * UPDATE * I'm going to leave this here in case anyone has a different solution, can explain how they do it, or why this works. This is what I figured out. In my system I use a 10th scale unit to pixels on non-retina displays and a 20th scale on retina displays. The iPhone is 640 pixels wide on retina and 320 pixels wide on non-retina (obsolete). So if I want something to be the full screen width I divide by 20 to get the OpenGL unit width. Then divide that by 2 to get the left and right unit position. Something 32 units wide centered on the screen goes from -16 to +16. Believe it or not I have an excel spreadsheet do all this math for me and output all the vertex data for my sprite sheet. It's an arbitrary thing I made up to do .1 units = 1 non-retina pixel or 2 retina pixels. I could have made it .01 units = 2 pixels and someday I might switch to that. But for now it's the other. So the width of the screen in units is 32.0, and that means the left most pixel is at -16.0 and the right most is at 16.0. After messing a bit I figured out that if I take the [0] value of an identity modelViewProjection matrix and multiply it by 16 I get the depth required to get 1:1 pixels. I don't know why. I don't know if the 16 is related to the screen size or just a lucky guess. But I did a test where I placed a sprite at that calculated depth and varied the FOV through all the valid values and the object stays steady on screen with 1:1 pixels. So now I'm just calculating the unityDepth that way. If someone gives me a better answer I'll checkmark it.

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• drawing thick, textured lines in OpenGL

  - by NateS

  I need to draw thick textured line segments in OpenGL. Actually I need curves made out of short line segments. Here is what I have: In the upper left is an example of two connected line segments. The second image shows once the lines are given width, they overlap. If I apply a texture that uses translucency, the overlap looks terrible. The third image shows that both lines are shortened by half the amount necessary to make the thick line corners just touch. This way I can fill the space between the lines with a triangle. On the right you can see this works well (ignore the horizontal line when the crappy texture repeats). But it doesn't always work well. In the bottom left the curve is made of many short line segments. Note the incorrect texture application. My program is written in Java, making use of the LWJGL OpenGL binding (and minor use of Slick, a 2D helper framework). I've made a zip file that contains an executable JAR so you can easily see the problem. It also has the Java code (there is only one source file) and an Eclipse project, so you can instantly run it through Eclipse and hack at it if you like. Here she is: http://n4te.com/temp/lines.zip To run, execute "java -jar lines.jar". You may need "-Djava.library.path=." before -jar if you are not on Windows. Press space to toggle texture/wireframe. The wireframe only shows the line segments, the triangle between them isn't drawn. I don't need to draw arbitrary lines, just bezier curves similar to what you see in the program. Sorry the code is a bit messy, once I have a solution I will refactor. I have investigated using GLUtessellator. It greatly simplified construction of the line, but I found that applying the texture was perfect. It worked most of the time (top image below), but long vertical curves would have severe texture distortion (bottom image below): This turned out to be much easier to code, but in the end worse than my approach. I believe what I'm trying to do is called "line tessellation" or "stroke tessellation". I assume this has been solved already? Is there standard code I can leverage? Otherwise, how can I fix my code so that the texture does not freak out on short, vertical curves?

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• Problem with waitable timers in Windows (timeSetEvent and CreateTimerQueueTimer)

  - by MusiGenesis

  I need high-resolution (more accurate than 1 millisecond) timing in my application. The waitable timers in Windows are (or can be made) accurate to the millisecond, but if I need a precise periodicity of, say, 35.7142857141 milliseconds, even a waitable timer with a 36 ms period will drift out of sync quickly. My "solution" to this problem (in ironic quotes because it's not working quite right) is to use a series of one-shot timers where I use the expiration of each timer to call the next timer. Normally a process like this would be subject to cumulative error over time, but in each timer callback I check the current time (with System.Diagnostics.Stopwatch) and use this to calculate what the period of the next timer needs to be (so if a timer happens to expire a little late, the next timer will automagically have a shorter period to compensate). This works as expected, except that after maybe 10-15 seconds the timer system seems to get bogged down, and a few timer callbacks here and there arrive anywhere from 25 to 100 milliseconds late. After a couple of seconds the problem goes away and everything runs smoothly again for 10-15 seconds, and then the stuttering again. Since I'm using Stopwatch to set each timer period, I'm also using it to monitor the arrival times of each timer callback. During the smooth-running periods, most (maybe 95%) of the intervals are either 35 or 36 milliseconds, and no intervals are ever more than 5 milliseconds away from the expected 35.7142857143. During the "glitchy" stretches, the distribution of intervals is very nearly identical, except that a very small number are unusually large (a couple more than 60 ms and one or two longer than 100 ms during maybe a 3-second stretch). This stuttering is very noticeable, and it's what I'm trying to fix, if possible. For the high-resolution timer, I was using the extremely antique timeSetEvent() multimedia timer from winmm.dll. In pursuit of this problem, I switched to using CreateTimerQueueTimer (along with timeBeginPeriod to set the high-resolution), but I'm seeing the same problem with both timer mechanisms. I've tried experimenting with the various flags for CreateTimerQueueTimer which determine which thread the timer runs on, but the stuttering appears no matter what. Is this just a fundamental problem with using timers in this way (i.e. using each one-shot timer to call the next)? If so, do I have any alternatives? One thing I was considering was to determine how many consecutive 1-millisecond-accuracy ticks would keep my within some arbitrary precision limit before I need to reset the timer. So, for example, if I wanted a 35.71428 period, I could let a 36 ms timer elapse 15 times before it was off by 5 milliseconds, then kill it and start a new one.

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• iPhone OpenGL ES: How do I use gravity vector to correctly transform scene for augmented reality

  - by gpdawson

  I'm trying figure out how to get an OpenGL specified object to be displayed correctly according to the device orientation (ie. according to the gravity vector from the accelerometer, and heading from compass). The GLGravity sample project has an example which is almost like this (despite ignoring heading), but it has some glitches. For example, the teapot jumps 180deg as the device viewing angle crosses the horizon, and it also rotates spuriously if you tilt the device from portrait into landscape. This is fine for the context of this app, as it just shows off an object and it doesn't matter that it does these things. But it means that the code just doesn't work when you attempt to emulate real life viewing of an OpenGL object according to the device's orientation. What happens is that it almost works, but the heading rotation you apply from the compass gets "corrupted" by the spurious additional rotations seen in the GLGravity example project. Can anyone provide sample code that shows how to adjust correctly for the device orientation (ie. gravity vector), or to fix the GLGravity example so that it doesn't include spurious heading changes? //Clear matrix to be used to rotate from the current referential to one based on the gravity vector bzero(matrix, sizeof(matrix)); matrix[3][3] = 1.0; //Setup first matrix column as gravity vector matrix[0][0] = accel[0] / length; matrix[0][1] = accel[1] / length; matrix[0][2] = accel[2] / length; //Setup second matrix column as an arbitrary vector in the plane perpendicular to the gravity vector {Gx, Gy, Gz} defined by by the equation "Gx * x + Gy * y + Gz * z = 0" in which we arbitrarily set x=0 and y=1 matrix[1][0] = 0.0; matrix[1][1] = 1.0; matrix[1][2] = -accel[1] / accel[2]; length = sqrtf(matrix[1][0] * matrix[1][0] + matrix[1][1] * matrix[1][1] + matrix[1][2] * matrix[1][2]); matrix[1][0] /= length; matrix[1][1] /= length; matrix[1][2] /= length; //Setup third matrix column as the cross product of the first two matrix[2][0] = matrix[0][1] * matrix[1][2] - matrix[0][2] * matrix[1][1]; matrix[2][1] = matrix[1][0] * matrix[0][2] - matrix[1][2] * matrix[0][0]; matrix[2][2] = matrix[0][0] * matrix[1][1] - matrix[0][1] * matrix[1][0]; //Finally load matrix glMultMatrixf((GLfloat*)matrix);

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• Prevent recursive CTE visiting nodes multiple times

  - by bacar

  Consider the following simple DAG: 1->2->3->4 And a table, #bar, describing this (I'm using SQL Server 2005): parent_id child_id 1 2 2 3 3 4 //... other edges, not connected to the subgraph above Now imagine that I have some other arbitrary criteria that select the first and last edges, i.e. 1-2 and 3-4. I want to use these to find the rest of my graph. I can write a recursive CTE as follows (I'm using terminology from MSDN): with foo(parent_id,child_id) as ( // anchor member that happens to select first and last edges: select parent_id,child_id from #bar where parent_id in (1,3) union all // recursive member: select #bar.* from #bar join foo on #bar.parent_id = foo.child_id ) select parent_id,child_id from foo However, this results in edge 3-4 being selected twice: parent_id child_id 1 2 3 4 2 3 3 4 // 2nd appearance! How can I prevent the query from recursing into subgraphs that have already been described? I could achieve this if, in my "recursive member" part of the query, I could reference all data that has been retrieved by the recursive CTE so far (and supply a predicate indicating in the recursive member excluding nodes already visited). However, I think I can access data that was returned by the last iteration of the recursive member only. This will not scale well when there is a lot of such repetition. Is there a way of preventing this unnecessary additional recursion? Note that I could use "select distinct" in the last line of my statement to achieve the desired results, but this seems to be applied after all the (repeated) recursion is done, so I don't think this is an ideal solution. Edit - hainstech suggests stopping recursion by adding a predicate to exclude recursing down paths that were explicitly in the starting set, i.e. recurse only where foo.child_id not in (1,3). That works for the case above only because it simple - all the repeated sections begin within the anchor set of nodes. It doesn't solve the general case where they may not be. e.g., consider adding edges 1-4 and 4-5 to the above set. Edge 4-5 will be captured twice, even with the suggested predicate. :(

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• How do I use the gravity vector to correctly transform scene for augmented reality?

  - by gpdawson

  I'm trying figure out how to get an OpenGL specified object to be displayed correctly according to the device orientation (ie. according to the gravity vector from the accelerometer, and heading from compass). The GLGravity sample project has an example which is almost like this (despite ignoring heading), but it has some glitches. For example, the teapot jumps 180deg as the device viewing angle crosses the horizon, and it also rotates spuriously if you tilt the device from portrait into landscape. This is fine for the context of this app, as it just shows off an object and it doesn't matter that it does these things. But it means that the code just doesn't work when you attempt to emulate real life viewing of an OpenGL object according to the device's orientation. What happens is that it almost works, but the heading rotation you apply from the compass gets "corrupted" by the spurious additional rotations seen in the GLGravity example project. Can anyone provide sample code that shows how to adjust correctly for the device orientation (ie. gravity vector), or to fix the GLGravity example so that it doesn't include spurious heading changes? //Clear matrix to be used to rotate from the current referential to one based on the gravity vector bzero(matrix, sizeof(matrix)); matrix[3][3] = 1.0; //Setup first matrix column as gravity vector matrix[0][0] = accel[0] / length; matrix[0][1] = accel[1] / length; matrix[0][2] = accel[2] / length; //Setup second matrix column as an arbitrary vector in the plane perpendicular to the gravity vector {Gx, Gy, Gz} defined by by the equation "Gx * x + Gy * y + Gz * z = 0" in which we arbitrarily set x=0 and y=1 matrix[1][0] = 0.0; matrix[1][1] = 1.0; matrix[1][2] = -accel[1] / accel[2]; length = sqrtf(matrix[1][0] * matrix[1][0] + matrix[1][1] * matrix[1][1] + matrix[1][2] * matrix[1][2]); matrix[1][0] /= length; matrix[1][1] /= length; matrix[1][2] /= length; //Setup third matrix column as the cross product of the first two matrix[2][0] = matrix[0][1] * matrix[1][2] - matrix[0][2] * matrix[1][1]; matrix[2][1] = matrix[1][0] * matrix[0][2] - matrix[1][2] * matrix[0][0]; matrix[2][2] = matrix[0][0] * matrix[1][1] - matrix[0][1] * matrix[1][0]; //Finally load matrix glMultMatrixf((GLfloat*)matrix);

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• How to setup named instances using StructureMap profiles?

  - by khaledh

  I've done quite a bit of googling and searching here on SO, but couldn't find a similar question or answer. In typical SM configuration you can add multiple named instances for a single PluginType: ForRequestedType<IFoo>() .AddInstances( x => { x.OfConcreteType<FooA>().WithName( "FooA" ); x.OfConcreteType<FooB>().WithName( "FooB" ); } ); No problem there. The problem is that I can't do the same when creating a profile. Most examples explaining how to use profiles use the For<>() method of the passed ProfileExpression: CreateProfile( "Default", p => { p.For<IFoo>().UseConcreteType<FooC>(); } ); I can't seem to find a way to add multiple named instances for the same PluginType as you can do above with regular configuration. The only other method available through ProfileExpression is Type<>(), but I'm not sure if it can be used for this purpose. Edit: I tried to use Type<>() instead of For<>() and it seems to be taking me in the right direction, but I bumped into another problem. To better explain it here's a better example of what I'm trying to do (this is what I posted to the structuremap-users group, no answer yet): ObjectFactory.Initialize( x => { x.CreateProfile( "Nissan", p => { p.Type<ICar>().Is.OfConcreteType<NewNissanCar>().WithName( "New" ); p.Type<ICar>().Is.OfConcreteType<OldNissanCar>().WithName( "Old" ); } ); x.CreateProfile( "Honda", p => { p.Type<ICar>().Is.OfConcreteType<NewHondaCar>().WithName( "New" ); p.Type<ICar>().Is.OfConcreteType<OldHondaCar>().WithName( "Old" ); } ); } ); ObjectFactory.Profile = "Nissan"; ICar newCar = ObjectFactory.GetNamedInstance<ICar>( "New" ); // -> returns NewHondaCar ICar car = ObjectFactory.GetInstance<ICar>(); // -> returns OldNissanCar So even though I set the profile to "Nissan", GetNamedInstance<>("New") returned an instance from the incorrect profile - it should've returned NewNissanCar instead of NewHondaCar. Interestingly, GetInstance<>() uses the correct profile, but because I can't pass an instance name, it returns an arbitrary concrete type from that profile that implements ICar (I guess it just returns the last concrete type added for that interface).

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• Server side Xforms form validation and integration into ASP.NET

  - by Nigel

  I have recently been investigating methods of creating web-based forms for an ASP.NET web application that can be edited and managed at runtime. For example an administrator might wish to add a new validation rule or a new set of fields. The holy grail would provide a means of specifying a form along with (potentially very complex) arbitrary validation rules, and allocation of data sources for each field. The specification would then be used to update the deployed form in the web application which would then validate submissions both on the client side and on the server side. My investigations led me to Xforms and a number of technologies that support it. One solution appears to be IBM Lotus Forms, but this requires a very large investment in terms of infrastructure, which makes it infeasible, although the forms designer may be useful as a stand-alone tool for creating the forms. I have also discounted browser plug-ins as the form must be publicly visible and cross-browser compliant. I have noticed that there are numerous javascript libraries that provide client side implementations given an Xforms schema. These would provide a partial solution but server side validation is still a requirement. Another option seems to involve the use of server side solutions such as the Java application Orbeon. Orbeon provides a tool for specifying the forms (although not as rich as Lotus Forms Designer), but the most interesting point is that it can translate an XForms schema into an XHTML form complete with validation. The fact that it is written in Java is not a big problem if it is possible to integrate with the existing ASP.NET application. So my question is whether anyone has done this before. It sounds like a problem that should have been solved but is inherently very complex. It seems possible to use an off-the-shelf tool to design the form and export it to an Xforms schema and xhtml form, and it seems possible to take that xforms schema and form and publish it using a client side library. What seems to be difficult is providing a means of validating the form submission on the server side and integrating the process nicely with .NET (although it seems the .NET community doesn't involve themselves with XForms; please correct me if I'm wrong on this count). I would be more than happy if a product provided something simple like a web service that could validate a submission against a schema. Maybe Orbeon does this but I'd be grateful if somebody in the know could point me in the right direction before I research it further. Many thanks.

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