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• Time required for a process to complete

  - by yelkawar

  I am new to C# world. I am attempting to calculate time taken by a algorithum for the purpose of comparison. Following code measures the elapsed time from when a subroutine is called until the subroutine returns to the main program.This example is taken from "Data structures through C#" by Michael McMillan. After running this program the output is Time=0, which is incorrect. The program appears to be logically correct. Can anybody help me. Following is the code using System.Collections.Generic; using System.Collections; using System.Linq; using System.Text; namespace Chap1 { class Program { static void Main(string[] args) { int num1 = 100; int num2 = 200; Console.WriteLine("num1: " + num1); Console.WriteLine("num2: " + num2); Swap<int>(ref num1, ref num2); Console.WriteLine("num1: " + num1); Console.WriteLine("num2: " + num2); string str1 = "Sam"; string str2 = "Tom"; Console.WriteLine("String 1: " + str1); Console.WriteLine("String 2: " + str2); Swap<string>(ref str1, ref str2); Console.WriteLine("String 1: " + str1); Console.WriteLine("String 2: " + str2); Console.ReadKey(); } static void Swap<T>(ref T val1, ref T val2) { T temp; temp = val1; val1 = val2; val2 = temp; } } class Timing { TimeSpan StartTiming; TimeSpan duration; public Timing() { StartTiming = new TimeSpan(0); duration = new TimeSpan(0); } public TimeSpan startTime() { GC.Collect(); GC.WaitForPendingFinalizers(); StartTiming = Process.GetCurrentProcess().Threads[0].UserProcessorTime; return StartTiming; } public void stopTime() { duration = Process.GetCurrentProcess().Threads[0].UserProcessorTime.Subtract(StartTiming); } public TimeSpan result() { return duration; } } }

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• AJAX response not valid in C++ but Apache

  - by fehergeri

  I want to make a server written in C++ to power my game. I learned the basics of sockets and wrote a basic chat program that worked well. Now I want to create an HTTP server like Apache, but only for the AJAX request-response part. I think just for the beginning i copied one Apache response text, and i sent the exact response with the C++ server program. The problem that is that the browser (Firefox) connnects to the apache and everything works fine, except all of the requests get a correct response. But if i send this with the C++ client, then FireBug tells me that the response status is OK (200) but there is no actual response text. (How is this possible?) This response-text is exactly the same what apache sends. I made a bit-bit comparison and they were the same. The php file wich is the original response <?php echo "AS";echo rand(0,9); ?> And the origional source code: Socket.h http://pastebin.com/bW9qxtrR Socket.cpp http://pastebin.com/S3c8RFM7 main.cpp http://pastebin.com/ckExuXsR index.html http://pastebin.com/mcfEEqPP < this is the requester file. ajax.js http://pastebin.com/uXJe9hVC benchmark.js http://pastebin.com/djSYtKg9 jQuery is not needed. The main.cpp there is lot of trash code like main3 and main4 functions, these do not affect the result. I know that the response stuff in the C++ code is not really good because the connection closing is not the best; I will fix that later now I want to send a success response first. UPDATE: now i tested today a lot again and i find out there is no problem with the socket. I used the fiddler program to capture the the good answer and to capture the bad. They were the same. After this i turned off my socket application, and forced fiddler to auto respond, and the answer from the 'bad' answer still bat. So after that i replaced the bad with the good and nothing happedned. The bad answer with the good text still bad on the :8888 port but the other on the original :80 port was good, but they were absolutly the same and the same program sended it (fiddler) i think there is something missing if the response is not on the same server address (even not the same port). UPDATE: oh my god! i cant send ajax request to a remote server. now i know this.

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• Are there known problems with >= and <= and the eval function in JS?

  - by Augier

  I am currently writing a JS rules engine which at one point needs to evaluate boolean expressions using the eval() function. Firstly I construct an equation as such: var equation = "relation.relatedTrigger.previousValue" + " " + relation.operator + " " + "relation.value"; relation.relatedTrigger.previousValue is the value I want to compare. relation.operator is the operator (either "==", "!=", <=, "<", "", ="). relation.value is the value I want to compare with. I then simply pass this string to the eval function and it returns true or false as such: return eval(equation); This works absolutely fine (with words and numbers) or all of the operators except for = and <=. E.g. When evaluating the equation: relation.relatedTrigger.previousValue <= 100 It returns true when previousValue = 0,1,10,100 & all negative numbers but false for everything in between. I would greatly appreciate the help of anyone to either answer my question or to help me find an alternative solution. Regards, Augier. P.S. I don't need a speech on the insecurities of the eval() function. Any value given to relation.relatedTrigger.previousValue is predefined. edit: Here is the full function: function evaluateRelation(relation) { console.log("Evaluating relation") var currentValue; //if multiple values if(relation.value.indexOf(";") != -1) { var values = relation.value.split(";"); for (x in values) { var equation = "relation.relatedTrigger.previousValue" + " " + relation.operator + " " + "values[x]"; currentValue = eval(equation); if (currentValue) return true; } return false; } //if single value else { //Evaluate the relation and get boolean var equation = "relation.relatedTrigger.previousValue" + " " + relation.operator + " " + "relation.value"; console.log("relation.relatedTrigger.previousValue " + relation.relatedTrigger.previousValue); console.log(equation); return eval(equation); } } Answer: Provided by KennyTM below. A string comparison doesn't work. Converting to a numerical was needed.

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• How to change quicksort to output elements in descending order?

  - by masato-san

  Hi, I wrote a quicksort algorithm however, I would like to make a change somewhere so that this quicksort would output elements in descending order. I searched and found that I can change the comparison operator (<) in partition() to other way around (like below). //This is snippet from partition() function while($array[$l] < $pivot) { $l++; } while($array[$r] > $pivot) { $r--; } But it is not working.. If I quicksort the array below, $array = (3,9,5,7); should be: $array = (9,7,5,3) But actual output is: $array = (3,5,7,9) Below is my quicksort which trying to output elements in descending order. How should I make change to sort in descending order? If you need any clarification please let me know. Thanks! $array = (3,9,5,7); $app = new QuicksortDescending(); $app->quicksort($array, 0, count($array)); print_r($array); class QuicksortDescending { public function partitionDesc(&$array, $left, $right) { $l = $left; $r = $right; $pivot = $array[($right+$left)/2]; while($l <= $r) { while($array[$l] > $pivot) { $l++; } while($array[$r] < $pivot) { $r--; } if($l <= $r) {// if L and R haven't cross $this->swap($array, $l, $r); $l ++; $j --; } } return $l; } public function quicksortDesc(&$array, $left, $right) { $index = $this->partition($array, $left, $right); if($left < $index-1) { //if there is more than 1 element to sort on right subarray $this->quicksortDesc($array, $left, $index-1); } if($index < $right) { //if there is more than 1 element to sort on right subarray $this->quicksortDesc($array, $index, $right); } } }

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• Setting up Netbeans/Eclipse for Linux Kernel Development

  - by red.october

  Hi: I'm doing some Linux kernel development, and I'm trying to use Netbeans. Despite declared support for Make-based C projects, I cannot create a fully functional Netbeans project. This is despite compiling having Netbeans analyze a kernel binary that was compiled with full debugging information. Problems include: files are wrongly excluded: Some files are incorrectly greyed out in the project, which means Netbeans does not believe they should be included in the project, when in fact they are compiled into the kernel. The main problem is that Netbeans will miss any definitions that exist in these files, such as data structures and functions, but also miss macro definitions. cannot find definitions: Pretty self-explanatory - often times, Netbeans cannot find the definition of something. This is partly a result of the above problem. can't find header files: self-explanatory I'm wondering if anyone has had success with setting up Netbeans for Linux kernel development, and if so, what settings they used. Ultimately, I'm looking for Netbeans to be able to either parse the Makefile (preferred) or extract the debug information from the binary (less desirable, since this can significantly slow down compilation), and automatically determine which files are actually compiled and which macros are actually defined. Then, based on this, I would like to be able to find the definitions of any data structure, variable, function, etc. and have complete auto-completion. Let me preface this question with some points: I'm not interested in solutions involving Vim/Emacs. I know some people like them, but I'm not one of them. As the title suggest, I would be also happy to know how to set-up Eclipse to do what I need While I would prefer perfect coverage, something that only misses one in a million definitions is obviously fine SO's useful "Related Questions" feature has informed me that the following question is related: http://stackoverflow.com/questions/149321/what-ide-would-be-good-for-linux-kernel-driver-development. Upon reading it, the question is more of a comparison between IDE's, whereas I'm looking for how to set-up a particular IDE. Even so, the user Wade Mealing seems to have some expertise in working with Eclipse on this kind of development, so I would certainly appreciate his (and of course all of your) answers. Cheers

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• SQL Server 2008: If Multiple Values Set In Other Mutliple Values Set

  - by AJH

  In SQL, is there anyway to accomplish something like this? This is based off a report built in SQL Server Report Builder, where the user can specify multiple text values as a single report parameter. The query for the report grabs all of the values the user selected and stores them in a single variable. I need a way for the query to return only records that have associations to EVERY value the user specified. -- Assume there's a table of Elements with thousands of entries. -- Now we declare a list of properties for those Elements to be associated with. create table #masterTable ( ElementId int, Text varchar(10) ) insert into #masterTable (ElementId, Text) values (1, 'Red'); insert into #masterTable (ElementId, Text) values (1, 'Coarse'); insert into #masterTable (ElementId, Text) values (1, 'Dense'); insert into #masterTable (ElementId, Text) values (2, 'Red'); insert into #masterTable (ElementId, Text) values (2, 'Smooth'); insert into #masterTable (ElementId, Text) values (2, 'Hollow'); -- Element 1 is Red, Coarse, and Dense. Element 2 is Red, Smooth, and Hollow. -- The real table is actually much much larger than this; this is just an example. -- This is me trying to replicate how SQL Server Report Builder treats -- report parameters in its queries. The user selects one, some, all, -- or no properties from a list. The written query treats the user's -- selections as a single variable called @Properties. -- Example scenario 1: User only wants to see Elements that are BOTH Red and Dense. select e.* from Elements e where (@Properties) --ideally a set containing only Red and Dense in (select Text from #masterTable where ElementId = e.Id) --ideally a set containing only Red, Coarse, and Dense --Both Red and Dense are within Element 1's properties (Red, Coarse, Dense), so Element 1 gets returned, but not Element 2. -- Example scenario 2: User only wants to see Elements that are BOTH Red and Hollow. select e.* from Elements e where (@Properties) --ideally a set containing only Red and Hollow in (select Text from #masterTable where ElementId = e.Id) --Both Red and Hollow are within Element 2's properties (Red, Smooth, Hollow), so Element 2 gets returned, but not Element 1. --Example Scenario 3: User only picked the Red option. select e.* from Elements e where (@Properties) --ideally a set containing only Red in (select Text from #masterTable where ElementId = e.Id) --Red is within both Element 1 and Element 2's properties, so both Element 1 and Element 2 get returned. The above syntax doesn't actually work because SQL doesn't seem to allow multiple values on the left side of the "in" comparison. Error that returns: Subquery returned more than 1 value. This is not permitted when the subquery follows =, !=, <, <= , >, >= or when the subquery is used as an expression. Am I even on the right track here? Sorry if the example looks long-winded or confusing.

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• GCC problem with raw double type comparisons

  - by Monomer

  I have the following bit of code, however when compiling it with GCC 4.4 with various optimization flags I get some unexpected results when its run. #include <iostream> int main() { const unsigned int cnt = 10; double lst[cnt] = { 0.0 }; const double v[4] = { 131.313, 737.373, 979.797, 731.137 }; for(unsigned int i = 0; i < cnt; ++i) { lst[i] = v[i % 4] * i; } for(unsigned int i = 0; i < cnt; ++i) { double d = v[i % 4] * i; if(lst[i] != d) { std::cout << "error @ : " << i << std::endl; return 1; } } return 0; } when compiled with: "g++ -pedantic -Wall -Werror -O1 -o test test.cpp" I get the following output: "error @ : 3" when compiled with: "g++ -pedantic -Wall -Werror -O2 -o test test.cpp" I get the following output: "error @ : 3" when compiled with: "g++ -pedantic -Wall -Werror -O3 -o test test.cpp" I get no errors when compiled with: "g++ -pedantic -Wall -Werror -o test test.cpp" I get no errors I do not believe this to be an issue related to rounding, or epsilon difference in the comparison. I've tried this with Intel v10 and MSVC 9.0 and they all seem to work as expected. I believe this should be nothing more than a bitwise compare. If I replace the if-statement with the following: if (static_cast<long long int>(lst[i]) != static_cast<long long int>(d)), and add "-Wno-long-long" I get no errors in any of the optimization modes when run. If I add std::cout << d << std::endl; before the "return 1", I get no errors in any of the optimization modes when run. Is this a bug in my code, or is there something wrong with GCC and the way it handles the double type?

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• Perl: Compare and edit underlying structure in hash

  - by Mahfuzur Rahman Pallab

  I have a hash of complex structure and I want to perform a search and replace. The first hash is like the following: $VAR1 = { abc => { 123 => ["xx", "yy", "zy"], 456 => ["ab", "cd", "ef"] }, def => { 659 => ["wx", "yg", "kl"], 456 => ["as", "sd", "df"] }, mno => { 987 => ["lk", "dm", "sd"] }, } and I want to iteratively search for all '123'/'456' elements, and if a match is found, I need to do a comparison of the sublayer, i.e. of ['ab','cd','ef'] and ['as','sd','df'] and in this case, keep only the one with ['ab','cd','ef']. So the output will be as follows: $VAR1 = { abc => { 123 => ["xx", "yy", "zy"], 456 => ["ab", "cd", "ef"] }, def => { 659 => ["wx", "yg", "kl"] }, mno => { 987 => ["lk", "dm", "sd"] }, } So the deletion is based on the substructure, and not index. How can it be done? Thanks for the help!! Lets assume that I will declare the values to be kept, i.e. I will keep 456 = ["ab", "cd", "ef"] based on a predeclared value of ["ab", "cd", "ef"] and delete any other instance of 456 anywhere else. The search has to be for every key. so the code will go through the hash, first taking 123 = ["xx", "yy", "zy"] and compare it against itself throughout the rest of the hash, if no match is found, do nothing. If a match is found, like in the case of 456 = ["ab", "cd", "ef"], it will compare the two, and as I have said that in case of a match the one with ["ab", "cd", "ef"] would be kept, it will keep 456 = ["ab", "cd", "ef"] and discard any other instances of 456 anywhere else in the hash, i.e. it will delete 456 = ["as", "sd", "df"] in this case.

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• Database schema for simple stats project

  - by Bubnoff

  Backdrop: I have a file hierarchy of cvs files for multiple locations named by dates they cover ...by month specifically. Each cvs file in the folder is named after the location. eg', folder name: 2010-feb contains: location1.csv location2.csv Each CSV file holds records like this: 2010-06-28, 20:30:00 , 0 2010-06-29, 08:30:00 , 0 2010-06-29, 09:30:00 , 0 2010-06-29, 10:30:00 , 0 2010-06-29, 11:30:00 , 0 meaning of record columns ( column names ): Date, time, # of sessions I have a perl script that pulls the data from this mess and originally I was going to store it as json files, but am thinking a database might be more appropriate long term ...comparing year to year trends ...fun stuff like that. Pt 2 - My question/problem: So I now have a REST service that coughs up json with a test database. My question is [ I suck at db design ], how best to design a database backend for this? I am thinking the following tables would suffice and keep it simple: Location: (PK)location_code, name session: (PK)id, (FK)location_code, month, hour, num_sessions I need to be able to average sessions (plus min and max) for each hour across days of week in addition to days of week in a given month or months. I've been using perl hashes to do this and am trying to decide how best to implement this with a database. Do you think stored procedures should be used? As to the database, depending on info gathered here, it will be postgresql or sqlite. If there is no compelling reason for postgresql I'll stick with sqlite. How and where should I compare the data to hours of operation. I am storing the hours of operation in a yaml file. I currently 'match' the hour in the data to a hash from the yaml to do this. Would a database open simpler methods? I am thinking I would do this comparison as I do now then insert the data. Can be recalled with: SELECT hour, num_sessions FROM session WHERE location_code=LOC1 Since only hours of operation are present, I do not need to worry about it. Should I calculate all results as I do now then store as a stats table for different 'reports'? This, rather than processing on demand? How would this look? Anyway ...I ramble. Thanks for reading! Bubnoff

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• Difference between Cloud and Virtualization

  - by Akash Kava

  Ops: This does not belong to ServerFault because it focuses on Programing Architecture. I have following questions regarding differences between Cloud and Virtualization.. How Cloud is different then Virtualization? Currently I tried to find out pricing of Rackspace, Amazone and all similar cloud providers, I found that our current 6 dedicated servers came cheaper then their pricing. So how one can claim cloud is cheaper? Is it cheaper only in comparison of normal hosting? We re organized our infrastructure in virtual environment to reduce or configuration overhead at time of failure, we did not have to rewrite any peice of code that is already written for earlier setup. So moving to virtualization does not require any re programming. But cloud is absoltely different and it will require entire reprogramming right? Is it really worth to recode when our current IT costs are 3-4 times lower then cloud hosting including raid backups and all sort of clustering for high availability? New programming architecture means new overheads of training staff, new methods of testing and new deployment schemes, does it justify over "on demand resource usage" words of cloud? We are having current development architecture with simple Server side ASP.NET WebServices with no local context and on client side Flex/Silverlight which offers pretty good REST architecture and its highly scalable. How does cloud differs from REST model of deployment? On storage, SQL Server or MySQL offers pretty good replication and high availibility then what is advantage in cloud? Data guarantee, one of our vendor hosting some other customer's app on cloud (one of most used), lost Entire Hard Disk (the virtual) and entire module in first 6 months. Second provider said its your duty to take backup, fine I agree, but no provider gives SLA for data guarantee, they give 99% uptime. However in most business apps, uptime is less important then data integrity. In our 10 years of dedicated hosting experience we had only one hard disk crash. This makes me little skeptical to go for cloud and loosing control over data. And I feel its just a big marketing buzz to sell virtulization in different form. Size of data, currently all providers charge very heavy for large data, if you are hosting only below 100GB cloud can be good alternative, but I think virtual servers and dedicated servers above 100GB to few TBs are still cheaper. Why would want to pay so high on cloud when there is no data guarentee as well as it doesnt say anything about redundancy. (I wish SO had something for spell check for Internet Explorer, sorry for wrong spellings in my post)

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• Code Golf: Countdown Number Game

  - by Noldorin

  Challenge Here is the task, inspired by the well-known British TV game show Countdown. The challenge should be pretty clear even without any knowledge of the game, but feel free to ask for clarifications. And if you fancy seeing a clip of this game in action, check out this YouTube clip. It features the wonderful late Richard Whitely in 1997. You are given 6 numbers, chosen at random from the set {1, 2, 3, 4, 5, 6, 8, 9, 10, 25, 50, 75, 100}, and a random target number between 100 and 999. The aim is to make use the six given numbers and the four common arithmetic operations (addition, subtraction, multiplication, division; all over the rational numbers) to generate the target - or as close as possible either side. Each number may only be used once at most, while each arithmetic operator may be used any number of times (including zero.) Note that it does not matter how many numbers are used. Write a function that takes the target number and set of 6 numbers (can be represented as list/collection/array/sequence) and returns the solution in any standard numerical notation (e.g. infix, prefix, postfix). The function must always return the closest-possible result to the target, and must run in at most 1 minute on a standard PC. Note that in the case where more than one solution exists, any single solution is sufficient. Examples: {50, 100, 4, 2, 2, 4}, target 203 e.g. 100 * 2 + 2 + (4 / 4) e.g. (100 + 50) * 4 * 2 / (4 + 2) {25, 4, 9, 2, 3, 10}, target 465 e.g. (25 + 10 - 4) * (9 * 2 - 3) {9, 8, 10, 5, 9, 7), target 241 e.g. ((10 + 9) * 9 * 7) + 8) / 5 Rules Other than mentioned in the problem statement, there are no further restrictions. You may write the function in any standard language (standard I/O is not necessary). The aim as always is to solve the task with the smallest number of characters of code. Saying that, I may not simply accept the answer with the shortest code. I'll also be looking at elegance of the code and time complexity of the algorithm! My Solution I'm attempting an F# solution when I find the free time - will post it here when I have something! Format Please post all answers in the following format for the purpose of easy comparison: Language Number of characters: ??? Fully obfuscated function: (code here) Clear (ideally commented) function: (code here) Any notes on the algorithm/clever shortcuts it takes.

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• Repeat Customers Each Year (Retention)

  - by spazzie

  I've been working on this and I don't think I'm doing it right. |D Our database doesn't keep track of how many customers we retain so we looked for an alternate method. It's outlined in this article. It suggests you have this table to fill in: Year Number of Customers Number of customers Retained in 2009 Percent (%) Retained in 2009 Number of customers Retained in 2010 Percent (%) Retained in 2010 .... 2008 2009 2010 2011 2012 Total The table would go out to 2012 in the headers. I'm just saving space. It tells you to find the total number of customers you had in your starting year. To do this, I used this query since our starting year is 2008: select YEAR(OrderDate) as 'Year', COUNT(distinct(billemail)) as Customers from dbo.tblOrder where OrderDate >= '2008-01-01' and OrderDate <= '2008-12-31' group by YEAR(OrderDate) At the moment we just differentiate our customers by email address. Then you have to search for the same names of customers who purchased again in later years (ours are 2009, 10, 11, and 12). I came up with this. It should find people who purchased in both 2008 and 2009. SELECT YEAR(OrderDate) as 'Year',COUNT(distinct(billemail)) as Customers FROM dbo.tblOrder o with (nolock) WHERE o.BillEmail IN (SELECT DISTINCT o1.BillEmail FROM dbo.tblOrder o1 with (nolock) WHERE o1.OrderDate BETWEEN '2008-1-1' AND '2009-1-1') AND o.BillEmail IN (SELECT DISTINCT o2.BillEmail FROM dbo.tblOrder o2 with (nolock) WHERE o2.OrderDate BETWEEN '2009-1-1' AND '2010-1-1') --AND o.OrderDate BETWEEN '2008-1-1' AND '2013-1-1' AND o.BillEmail NOT LIKE '%@halloweencostumes.com' AND o.BillEmail NOT LIKE '' GROUP BY YEAR(OrderDate) So I'm just finding the customers who purchased in both those years. And then I'm doing an independent query to find those who purchased in 2008 and 2010, then 08 and 11, and then 08 and 12. This one finds 2008 and 2010 purchasers: SELECT YEAR(OrderDate) as 'Year',COUNT(distinct(billemail)) as Customers FROM dbo.tblOrder o with (nolock) WHERE o.BillEmail IN (SELECT DISTINCT o1.BillEmail FROM dbo.tblOrder o1 with (nolock) WHERE o1.OrderDate BETWEEN '2008-1-1' AND '2009-1-1') AND o.BillEmail IN (SELECT DISTINCT o2.BillEmail FROM dbo.tblOrder o2 with (nolock) WHERE o2.OrderDate BETWEEN '2010-1-1' AND '2011-1-1') --AND o.OrderDate BETWEEN '2008-1-1' AND '2013-1-1' AND o.BillEmail NOT LIKE '%@halloweencostumes.com' AND o.BillEmail NOT LIKE '' GROUP BY YEAR(OrderDate) So you see I have a different query for each year comparison. They're all unrelated. So in the end I'm just finding people who bought in 2008 and 2009, and then a potentially different group that bought in 2008 and 2010, and so on. For this to be accurate, do I have to use the same grouping of 2008 buyers each time? So they bought in 2009 and 2010 and 2011, and 2012? This is where I'm worried and not sure how to proceed or even find such data. Any advice would be appreciated! Thanks!

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• Javascript: Writing a firefox extension with sockets

  - by Johnny Grass

  I need to write a firefox extension that creates a server socket (I think that's what it's called) and returns the browser's current url when a client application (running on the same computer) sends it a request. The thing is that I have no Java/Javascript background at all and I'm pressed for time so I am trying to hack something together from code samples. So far I've been mildly successful. I've been working with code from this question which is used in the open source Firefox exension PolyChrome I have the following code: var reader = { onInputStreamReady : function(input) { var input_stream = Components.classes["@mozilla.org/scriptableinputstream;1"] .createInstance(Components.interfaces.nsIScriptableInputStream); input_stream.init(input); input_stream.available(); var request = ''; while (input_stream.available()) { request = request + input_stream.read(512); } var checkString = "foo" if (request.toString() == checkString.toString()) { output_console('URL: ' + content.location.href); } else output_console("nothing"); var thread_manager = Components.classes["@mozilla.org/thread-manager;1"].getService(); input.asyncWait(reader,0,0,thread_manager.mainThread); } } var listener = { onSocketAccepted: function(serverSocket, clientSocket) { output_console("Accepted connection on "+clientSocket.host+":"+clientSocket.port); input = clientSocket.openInputStream(0, 0, 0).QueryInterface(Components.interfaces.nsIAsyncInputStream); output = clientSocket.openOutputStream(Components.interfaces.nsITransport.OPEN_BLOCKING, 0, 0); var thread_manager = Components.classes["@mozilla.org/thread-manager;1"].getService(); input.asyncWait(reader,0,0,thread_manager.mainThread); } } var serverSocket = Components.classes["@mozilla.org/network/server-socket;1"]. createInstance(Components.interfaces.nsIServerSocket); serverSocket.init(9999, true, 5); output_console("Opened socket on " + serverSocket.port); serverSocket.asyncListen(listener); I have a few questions. So far I can telnet into localhost and get a response, but my string comparison in the reader seems to fail even if I enter "foo". I don't get why. What am I missing? The sample code I'm using opens up a console window and prints output when I telnet into localhost. Ideally I would like the output to be returned as a response when the client sends a request to the server socket with a passphrase. How do I go about doing that? Is doing this a good idea? Does it create security vulnerabilities on the computer? How can I block connections to the socket from other computers? What is a good place to read about javascript sockets? My google searches have been pretty fruitless but then maybe I'm not using the right keywords.

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• MySQL search for user and their roles

  - by Jenkz

  I am re-writing the SQL which lets a user search for any other user on our site and also shows their roles. An an example, roles can be "Writer", "Editor", "Publisher". Each role links a User to a Publication. Users can take multiple roles within multiple publications. Example table setup: "users" : user_id, firstname, lastname "publications" : publication_id, name "link_writers" : user_id, publication_id "link_editors" : user_id, publication_id Current psuedo SQL: SELECT * FROM ( (SELECT user_id FROM users WHERE firstname LIKE '%Jenkz%') UNION (SELECT user_id FROM users WHERE lastname LIKE '%Jenkz%') ) AS dt JOIN (ROLES STATEMENT) AS roles ON roles.user_id = dt.user_id At the moment my roles statement is: SELECT dt2.user_id, dt2.publication_id, dt.role FROM ( (SELECT 'writer' AS role, link_writers.user_id, link_writers.publication_id FROM link_writers) UNION (SELECT 'editor' AS role, link_editors.user_id, link_editors.publication_id FROM link_editors) ) AS dt2 The reason for wrapping the roles statement in UNION clauses is that some roles are more complex and require a table join to find the publication_id and user_id. As an example "publishers" might be linked accross two tables "link_publishers": user_id, publisher_group_id "link_publisher_groups": publisher_group_id, publication_id So in that instance, the query forming part of my UNION would be: SELECT 'publisher' AS role, link_publishers.user_id, link_publisher_groups.publication_id FROM link_publishers JOIN link_publisher_groups ON lpg.group_id = lp.group_id I'm pretty confident that my table setup is good (I was warned off the one-table-for-all system when researching the layout). My problem is that there are now 100,000 rows in the users table and upto 70,000 rows in each of the link tables. Initial lookup in the users table is fast, but the joining really slows things down. How can I only join on the relevant roles? -------------------------- EDIT ---------------------------------- Explain above (open in a new window to see full resolution). The bottom bit in red, is the "WHERE firstname LIKE '%Jenkz%'" the third row searches WHERE CONCAT(firstname, ' ', lastname) LIKE '%Jenkz%'. Hence the large row count, but I think this is unavoidable, unless there is a way to put an index accross concatenated fields? The green bit at the top just shows the total rows scanned from the ROLES STATEMENT. You can then see each individual UNION clause (#6 - #12) which all show a large number of rows. Some of the indexes are normal, some are unique. It seems that MySQL isn't optimizing to use the dt.user_id as a comparison for the internal of the UNION statements. Is there any way to force this behaviour? Please note that my real setup is not publications and writers but "webmasters", "players", "teams" etc.

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• making a password-only auth with bcrypt and mongoose

  - by user3081123

  I want to create service that let you login only with password. You type a password and if this password exists - you are logged in and if it's not - username is generated and password is encrypted. I'm having some misunderstandings and hope someone would help me to show where I'm mistaken. I guess, it would look somewhat like this in agularjs First we receive a password in login controller. $scope.signup = function() { var user = { password: $scope.password, }; $http.post('/auth/signup', user); }; Send it via http.post and get in in our node server file. We are provided with a compare password bcrypt function userSchema.methods.comparePassword = function(candidatePassword, cb) { bcrypt.compare(candidatePassword, this.password, function(err, isMatch) { if (err) return cb(err); cb(null, isMatch); }); }; So right now we are creating function to catch our http request app.post('/auth/signup', function(req, res, next) { Inside we use a compair password function to realize if such password exists or not yet. So we have to encrypt a password with bcrypt to make a comparison First we hash it same way as in .pre var encPass; bcrypt.genSalt(10, function(err, salt) { if (err) return next(err); bcrypt.hash(req.body.password, salt, function(err, hash) { if (err) return next(err); encPass=hash; )}; )}; We have encrypted password stored in encPass so now we follow to finding a user in database with this password User.findOne({ password: encPass }, function(err, user) { if (user) { //user exists, it means we should pass an ID of this user to a controller to display it in a view. I don't know how. res.send({user.name}) //like this? How should controller receive this? With $http.post? } else { and now if user doesn't exist - we should create it with user ID generated by my function var nUser = new User({ name: generId(), password: req.body.password }); nUser.save(function(err) { if (err) return next(err); )}; )}; )}; Am I doing anything right? I'm pretty new to js and angular. If so - how do I throw a username back at controller? If someone is interested - this service exists for 100+ symbol passphrases so possibility of entering same passphrase as someone else is miserable. And yeah, If someone logged in under 123 password - the other guy will log in as same user if he entered 123 password, but hey, you are warned to make a big passphrase. So I'm confident about the idea and I only need a help with understanding and realization.

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• What's causing this background-image to display "incorrectly" in Opera and Firefox?

  - by Sukasa

  I know this is something I'm probably doing wrong, so please don't incinerate me for the thread title. I'm trying to put together a small personal website using HTML 5/CSS3. I've checked with the w3c validator and the site and CSS file fully conform according to the validator (However the validator has a warning attached that it might not be perfect). I'm not sure how to explain it without a picture, so here's a comparison of Chrome/Opera/Firefox: So, you can sorta see how in Chrome the background image is in one non-repeating piece, whereas in Opera/Firefox the image has, oddly, been broken up and placed slightly differently. I'm confident this is due to an error on my part, but I've had no luck at all figuring out why the image is being mangled in Opera and Firefox. Here's the CSS that's relevant to this issue: /* Content Pane */ .content { position: absolute; left: 220px; width: 800px; top: 80px; min-height: 550px; background-color: rgba(8,12,42,0.85); } /* Headers */ .content hgroup { background: url("Header_Flat.png") no-repeat left top; min-height: 38px; padding-left: 28px; text-shadow: 0 0 8px #FFA9FF; color: Black; text-decoration: none; } .content hgroup h1 { display: block; } .content hgroup h3 { display: inline; position: relative; top: -12px; left: 20px; text-shadow: 0 0 6px #AFF9FF; } .content hgroup h4 { display: inline; position: relative; top: -12px; left: 20px; font-size: xx-small; text-shadow: 0 0 6px #AFF9FF; } And the HTML: <hgroup> <h1>New Site!</h1> <h3>Now with Bloom!</h3> <h4> - Posted Tuesday, May 11th 2010</h4> </hgroup> Can anyone see what I'm doing wrong?

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• Compare array in loop

  - by user3626084

  I have 2 arrays with different sizes, in some cases one array can have more elements than the other array. However, I always need to compare the arrays using the same id. I need to get the other value with the same id in the other array I have tried this, but the problem happens when I compare the two arrays in a loop when the other array has more elements than one, because duplicate the loop and data , and it does not work. Here is what I've tried: <?php /// Actual Data Arrays /// $data_1=array("a1-fruits","b1-apple","c1-banana","d1-chocolate","e1-pear"); $data_2=array("b1-cars","e1-eggs"); /// for ($i=0;$i<count($data_1);$i++) { /// Explode ID $data_1 /// $exp_id=explode("-",$data_1[$i]); /// for ($h=0;$h<count($data_2);$h++) { /// Explode ID $data_2 /// $exp_id2=explode("-",$data_2[$h]); /// if ($exp_id[0]=="".$exp_id2[0]."") { print "".$data_2[$h].""; print "<br>"; } else { print "".$data_1[$i].""; print "<br>"; } /// } /// } ?> I want the following values : "a1-fruits" "b1-cars" "c1-banana" "d1-chocolate" "e1-eggs" Yet, I get this (which isn't what I want): a1-fruits a1-fruits b1-cars b1-apple c1-banana c1-banana d1-chocolate d1-chocolate e1-pear e1-eggs I tried everything I know and try to understand how I can do this because I don't understand how to compare these two arrays. The other problem is when one size has more elements than the other, the comparison always gives an error. I FIND THE SOLUTION TO THIS AND WORKING IN ALL : <?php /// Actual Data Arrays /// $data_1=array("a1-fruits","b1-apple","c1-banana","d1-chocolate","e1-pear"); $data_2=array("b1-cars","e1-eggs","d1-chocolate2"); /// for ($i=0;$i<count($data_1);$i++) { $show="bad"; /// Explode ID $data_1 /// $exp_id=explode("-",$data_1[$i]); /// for ($h=0;$h<count($data_2);$h++) { /// Explode ID $data_2 /// $exp_id2=explode("-",$data_2[$h]); /// if ($exp_id2[0]=="".$exp_id[0]."") { $show="ok"; print "".$data_2[$h]."<br>"; } /// } if ($show=="bad") { print "".$data_1[$i].""; print "<br>"; } /// } ?>

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• Numpy/Python performing terribly vs. Matlab

  - by Nissl

  Novice programmer here. I'm writing a program that analyzes the relative spatial locations of points (cells). The program gets boundaries and cell type off an array with the x coordinate in column 1, y coordinate in column 2, and cell type in column 3. It then checks each cell for cell type and appropriate distance from the bounds. If it passes, it then calculates its distance from each other cell in the array and if the distance is within a specified analysis range it adds it to an output array at that distance. My cell marking program is in wxpython so I was hoping to develop this program in python as well and eventually stick it into the GUI. Unfortunately right now python takes ~20 seconds to run the core loop on my machine while MATLAB can do ~15 loops/second. Since I'm planning on doing 1000 loops (with a randomized comparison condition) on ~30 cases times several exploratory analysis types this is not a trivial difference. I tried running a profiler and array calls are 1/4 of the time, almost all of the rest is unspecified loop time. Here is the python code for the main loop: for basecell in range (0, cellnumber-1): if firstcelltype == np.array((cellrecord[basecell,2])): xloc=np.array((cellrecord[basecell,0])) yloc=np.array((cellrecord[basecell,1])) xedgedist=(xbound-xloc) yedgedist=(ybound-yloc) if xloc>excludedist and xedgedist>excludedist and yloc>excludedist and yedgedist>excludedist: for comparecell in range (0, cellnumber-1): if secondcelltype==np.array((cellrecord[comparecell,2])): xcomploc=np.array((cellrecord[comparecell,0])) ycomploc=np.array((cellrecord[comparecell,1])) dist=math.sqrt((xcomploc-xloc)**2+(ycomploc-yloc)**2) dist=round(dist) if dist>=1 and dist<=analysisdist: arraytarget=round(dist*analysisdist/intervalnumber) addone=np.array((spatialraw[arraytarget-1])) addone=addone+1 targetcell=arraytarget-1 np.put(spatialraw,[targetcell,targetcell],addone) Here is the matlab code for the main loop: for basecell = 1:cellnumber; if firstcelltype==cellrecord(basecell,3); xloc=cellrecord(basecell,1); yloc=cellrecord(basecell,2); xedgedist=(xbound-xloc); yedgedist=(ybound-yloc); if (xloc>excludedist) && (yloc>excludedist) && (xedgedist>excludedist) && (yedgedist>excludedist); for comparecell = 1:cellnumber; if secondcelltype==cellrecord(comparecell,3); xcomploc=cellrecord(comparecell,1); ycomploc=cellrecord(comparecell,2); dist=sqrt((xcomploc-xloc)^2+(ycomploc-yloc)^2); if (dist>=1) && (dist<=100.4999); arraytarget=round(dist*analysisdist/intervalnumber); spatialsum(1,arraytarget)=spatialsum(1,arraytarget)+1; end end end end end end Thanks!

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• Why Swift is 100 times slower than C in this image processing test?

  - by xiaobai

  Like many other developers I have been very excited at the new Swift language from Apple. Apple has boasted its speed is faster than Objective C and can be used to write operating system. And from what I learned so far, it's a very type-safe language and able to have precisely control over the exact data type (like integer length). So it does look like having good potential handling performance critical tasks, like image processing, right? That's what I thought before I carried out a quick test. The result really surprised me. Here is a much simplified image alpha blending code snippet in C: test.c: #include <stdio.h> #include <stdint.h> #include <string.h> uint8_t pixels[640*480]; uint8_t alpha[640*480]; uint8_t blended[640*480]; void blend(uint8_t* px, uint8_t* al, uint8_t* result, int size) { for(int i=0; i<size; i++) { result[i] = (uint8_t)(((uint16_t)px[i]) *al[i] /255); } } int main(void) { memset(pixels, 128, 640*480); memset(alpha, 128, 640*480); memset(blended, 255, 640*480); // Test 10 frames for(int i=0; i<10; i++) { blend(pixels, alpha, blended, 640*480); } return 0; } I compiled it on my Macbook Air 2011 with the following command: gcc -O3 test.c -o test The 10 frame processing time is about 0.01s. In other words, it takes the C code 1ms to process one frame: $ time ./test real 0m0.010s user 0m0.006s sys 0m0.003s Then I have a Swift version of the same code: test.swift: let pixels = UInt8[](count: 640*480, repeatedValue: 128) let alpha = UInt8[](count: 640*480, repeatedValue: 128) let blended = UInt8[](count: 640*480, repeatedValue: 255) func blend(px: UInt8[], al: UInt8[], result: UInt8[], size: Int) { for(var i=0; i<size; i++) { var b = (UInt16)(px[i]) * (UInt16)(al[i]) result[i] = (UInt8)(b/255) } } for i in 0..10 { blend(pixels, alpha, blended, 640*480) } The build command line is: xcrun swift -O3 test.swift -o test Here I use the same O3 level optimization flag to make the comparison hopefully fair. However, the resulting speed is 100 time slower: $ time ./test real 0m1.172s user 0m1.146s sys 0m0.006s In other words, it takes Swift ~120ms to processing one frame which takes C just 1 ms. I also verified the memory initialization time in both test code are very small compared to the blend processing function time. What happened?

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• Multi-tenant ASP.NET MVC – Introduction

  - by zowens

  I’ve read a few different blogs that talk about multi-tenancy and how to resolve some of the issues surrounding multi-tenancy. What I’ve come to realize is that these implementations overcomplicate the issues and give only a muddy implementation! I’ve seen some really illogical code out there. I have recently been building a multi-tenancy framework for internal use at eagleenvision.net. Through this process, I’ve realized a few different techniques to make building multi-tenant applications actually quite easy. I will be posting a few different entries over the issue and my personal implementation. In this first post, I will discuss what multi-tenancy means and how my implementation will be structured.   So what’s the problem? Here’s the deal. Multi-tenancy is basically a technique of code-reuse of web application code. A multi-tenant application is an application that runs a single instance for multiple clients. Here the “client” is different URL bindings on IIS using ASP.NET MVC. The problem with different instances of the, essentially, same application is that you have to spin up different instances of ASP.NET. As the number of running instances of ASP.NET grows, so does the memory footprint of IIS. Stack Exchange shifted its architecture to multi-tenancy March. As the blog post explains, multi-tenancy saves cost in terms of memory utilization and physical disc storage. If you use the same code base for many applications, multi-tenancy just makes sense. You’ll reduce the amount of work it takes to synchronize the site implementations and you’ll thank your lucky stars later for choosing to use one application for multiple sites. Multi-tenancy allows the freedom of extensibility while relying on some pre-built code.   You’d think this would be simple. I have actually seen a real lack of reference material on the subject in terms of ASP.NET MVC. This is somewhat surprising given the number of users of ASP.NET MVC. However, I will certainly fill the void ;). Implementing a multi-tenant application takes a little thinking. It’s not straight-forward because the possibilities of implementation are endless. I have yet to see a great implementation of a multi-tenant MVC application. The only one that comes close to what I have in mind is Rob Ashton’s implementation (all the entries are listed on this page). There’s some really nasty code in there… something I’d really like to avoid. He has also written a library (MvcEx) that attempts to aid multi-tenant development. This code is even worse, in my honest opinion. Once I start seeing Reflection.Emit, I have to assume the worst :) In all seriousness, if his implementation makes sense to you, use it! It’s a fine implementation that should be given a look. At least look at the code. I will reference MvcEx going forward as a comparison to my implementation. I will explain why my approach differs from MvcEx and how it is better or worse (hopefully better).   Core Goals of my Multi-Tenant Implementation The first, and foremost, goal is to use Inversion of Control containers to my advantage. As you will see throughout this series, I pass around containers quite frequently and rely on their use heavily. I will be using StructureMap in my implementation. However, you could probably use your favorite IoC tool instead. <RANT> However, please don’t be stupid and abstract your IoC tool. Each IoC is powerful and by abstracting the capabilities, you’re doing yourself a real disservice. Who in the world swaps out IoC tools…? No one!</RANT> (It had to be said.) I will outline some of the goodness of StructureMap as we go along. This is really an invaluable tool in my tool belt and simple to use in my multi-tenant implementation. The second core goal is to represent a tenant as easily as possible. Just as a dependency container will be a first-class citizen, so will a tenant. This allows us to easily extend and use tenants. This will also allow different ways of “plugging in” tenants into your application. In my implementation, there will be a single dependency container for a single tenant. This will enable isolation of the dependencies of the tenant. The third goal is to use composition as a means to delegate “core” functions out to the tenant. More on this later.   Features In MvcExt, “Modules” are a code element of the infrastructure. I have simplified this concept and have named this “Features”. A feature is a simple element of an application. Controllers can be specified to have a feature and actions can have “sub features”. Each tenant can select features it needs and the other features will be hidden to the tenant’s users. My implementation doesn’t require something to be a feature. A controller can be common to all tenants. For example, (as you will see) I have a “Content” controller that will return the CSS, Javascript and Images for a tenant. This is common logic to all tenants and shouldn’t be hidden or considered a “feature”; Content is a core component.   Up next My next post will be all about the code. I will reveal some of the foundation to the way I do multi-tenancy. I will have posts dedicated to Foundation, Controllers, Views, Caching, Content and how to setup the tenants. Each post will be in-depth about the issues and implementation details, while adhering to my core goals outlined in this post. As always, comment with questions of DM me on twitter or send me an email.

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• Ajax Control Toolkit and Superexpert

  - by Stephen Walther

  Microsoft has asked my company, Superexpert Consulting, to take ownership of the development and maintenance of the Ajax Control Toolkit moving forward. In this blog entry, I discuss our strategy for improving the Ajax Control Toolkit. Why the Ajax Control Toolkit? The Ajax Control Toolkit is one of the most popular projects on CodePlex. In fact, some have argued that it is among the most successful open-source projects of all time. It consistently receives over 3,500 downloads a day (not weekends -- workdays). A mind-boggling number of developers use the Ajax Control Toolkit in their ASP.NET Web Forms applications. Why does the Ajax Control Toolkit continue to be such a popular project? The Ajax Control Toolkit fills a strong need in the ASP.NET Web Forms world. The Toolkit enables Web Forms developers to build richly interactive JavaScript applications without writing any JavaScript. For example, by taking advantage of the Ajax Control Toolkit, a Web Forms developer can add modal dialogs, popup calendars, and client tabs to a web application simply by dragging web controls onto a page. The Ajax Control Toolkit is not for everyone. If you are comfortable writing JavaScript then I recommend that you investigate using jQuery plugins instead of the Ajax Control Toolkit. However, if you are a Web Forms developer and you don’t want to get your hands dirty writing JavaScript, then the Ajax Control Toolkit is a great solution. The Ajax Control Toolkit is Vast The Ajax Control Toolkit consists of 40 controls. That’s a lot of controls (For the sake of comparison, jQuery UI consists of only 8 controls – those slackers J). Furthermore, developers expect the Ajax Control Toolkit to work on browsers both old and new. For example, people expect the Ajax Control Toolkit to work with Internet Explorer 6 and Internet Explorer 9 and every version of Internet Explorer in between. People also expect the Ajax Control Toolkit to work on the latest versions of Mozilla Firefox, Apple Safari, and Google Chrome. And, people expect the Ajax Control Toolkit to work with different operating systems. Yikes, that is a lot of combinations. The biggest challenge which my company faces in supporting the Ajax Control Toolkit is ensuring that the Ajax Control Toolkit works across all of these different browsers and operating systems. Testing, Testing, Testing Because we wanted to ensure that we could easily test the Ajax Control Toolkit with different browsers, the very first thing that we did was to set up a dedicated testing server. The dedicated server -- named Schizo -- hosts 4 virtual machines so that we can run Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, and Internet Explorer 9 at the same time (We also use the virtual machines to host the latest versions of Firefox, Chrome, Opera, and Safari). The five developers on our team (plus me) can each publish to a separate FTP website on the testing server. That way, we can quickly test how changes to the Ajax Control Toolkit affect different browsers. QUnit Tests for the Ajax Control Toolkit Introducing regressions – introducing new bugs when trying to fix existing bugs – is the concern which prevents me from sleeping well at night. There are so many people using the Ajax Control Toolkit in so many unique scenarios, that it is difficult to make improvements to the Ajax Control Toolkit without introducing regressions. In order to avoid regressions, we decided early on that it was extremely important to build good test coverage for the 40 controls in the Ajax Control Toolkit. We’ve been focusing a lot of energy on building automated JavaScript unit tests which we can use to help us discover regressions. We decided to write the unit tests with the QUnit test framework. We picked QUnit because it is quickly becoming the standard unit testing framework in the JavaScript world. For example, it is the unit testing framework used by the jQuery team, the jQuery UI team, and many jQuery UI plugin developers. We had to make several enhancements to the QUnit framework in order to test the Ajax Control Toolkit. For example, QUnit does not support tests which include postbacks. We modified the QUnit framework so that it works with IFrames so we could perform postbacks in our automated tests. At this point, we have written hundreds of QUnit tests. For example, we have written 135 QUnit tests for the Accordion control. The QUnit tests are included with the Ajax Control Toolkit source code in a project named AjaxControlToolkit.Tests. You can run all of the QUnit tests contained in the project by opening the Default.aspx page. Automating the QUnit Tests across Multiple Browsers Automated tests are useless if no one ever runs them. In order for the QUnit tests to be useful, we needed an easy way to run the tests automatically against a matrix of browsers. We wanted to run the unit tests against Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, Internet Explorer 9, Firefox, Chrome, and Safari automatically. Expecting a developer to run QUnit tests against every browser after every check-in is just too much to expect. It takes 20 seconds to run the Accordion QUnit tests. We are testing against 8 browsers. That would require the developer to open 8 browsers and wait for the results after each change in code. Too much work. Therefore, we built a JavaScript Test Server. Our JavaScript Test Server project was inspired by John Resig’s TestSwarm project. The JavaScript Test Server runs our QUnit tests in a swarm of browsers (running on different operating systems) automatically. Here’s how the JavaScript Test Server works: 1. We created an ASP.NET page named RunTest.aspx that constantly polls the JavaScript Test Server for a new set of QUnit tests to run. After the RunTest.aspx page runs the QUnit tests, the RunTest.aspx records the test results back to the JavaScript Test Server. 2. We opened the RunTest.aspx page on instances of Internet Explorer 6, Internet Explorer 7, Internet Explorer 8, Internet Explorer 9, FireFox, Chrome, Opera, Google, and Safari. Now that we have the JavaScript Test Server setup, we can run all of our QUnit tests against all of the browsers which we need to support with a single click of a button. A New Release of the Ajax Control Toolkit Each Month The Ajax Control Toolkit Issue Tracker contains over one thousand five hundred open issues and feature requests. So we have plenty of work on our plates J At CodePlex, anyone can vote for an issue to be fixed. Originally, we planned to fix issues in order of their votes. However, we quickly discovered that this approach was inefficient. Constantly switching back and forth between different controls was too time-consuming. It takes time to re-familiarize yourself with a control. Instead, we decided to focus on two or three controls each month and really focus on fixing the issues with those controls. This way, we can fix sets of related issues and avoid the randomization caused by context switching. Our team works in monthly sprints. We plan to do another release of the Ajax Control Toolkit each and every month. So far, we have competed one release of the Ajax Control Toolkit which was released on April 1, 2011. We plan to release a new version in early May. Conclusion Fortunately, I work with a team of smart developers. We currently have 5 developers working on the Ajax Control Toolkit (not full-time, they are also building two very cool ASP.NET MVC applications). All the developers who work on our team are required to have strong JavaScript, jQuery, and ASP.NET MVC skills. In the interest of being as transparent as possible about our work on the Ajax Control Toolkit, I plan to blog frequently about our team’s ongoing work. In my next blog entry, I plan to write about the two Ajax Control Toolkit controls which are the focus of our work for next release.

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• C# 4: The Curious ConcurrentDictionary

  - by James Michael Hare

  In my previous post (here) I did a comparison of the new ConcurrentQueue versus the old standard of a System.Collections.Generic Queue with simple locking.  The results were exactly what I would have hoped, that the ConcurrentQueue was faster with multi-threading for most all situations.  In addition, concurrent collections have the added benefit that you can enumerate them even if they're being modified. So I set out to see what the improvements would be for the ConcurrentDictionary, would it have the same performance benefits as the ConcurrentQueue did?  Well, after running some tests and multiple tweaks and tunes, I have good and bad news. But first, let's look at the tests.  Obviously there's many things we can do with a dictionary.  One of the most notable uses, of course, in a multi-threaded environment is for a small, local in-memory cache.  So I set about to do a very simple simulation of a cache where I would create a test class that I'll just call an Accessor.  This accessor will attempt to look up a key in the dictionary, and if the key exists, it stops (i.e. a cache "hit").  However, if the lookup fails, it will then try to add the key and value to the dictionary (i.e. a cache "miss").  So here's the Accessor that will run the tests: 1: internal class Accessor 2: { 3: public int Hits { get; set; } 4: public int Misses { get; set; } 5: public Func<int, string> GetDelegate { get; set; } 6: public Action<int, string> AddDelegate { get; set; } 7: public int Iterations { get; set; } 8: public int MaxRange { get; set; } 9: public int Seed { get; set; } 10:  11: public void Access() 12: { 13: var randomGenerator = new Random(Seed); 14:  15: for (int i=0; i<Iterations; i++) 16: { 17: // give a wide spread so will have some duplicates and some unique 18: var target = randomGenerator.Next(1, MaxRange); 19:  20: // attempt to grab the item from the cache 21: var result = GetDelegate(target); 22:  23: // if the item doesn't exist, add it 24: if(result == null) 25: { 26: AddDelegate(target, target.ToString()); 27: Misses++; 28: } 29: else 30: { 31: Hits++; 32: } 33: } 34: } 35: } Note that so I could test different implementations, I defined a GetDelegate and AddDelegate that will call the appropriate dictionary methods to add or retrieve items in the cache using various techniques. So let's examine the three techniques I decided to test: Dictionary with mutex - Just your standard generic Dictionary with a simple lock construct on an internal object. Dictionary with ReaderWriterLockSlim - Same Dictionary, but now using a lock designed to let multiple readers access simultaneously and then locked when a writer needs access. ConcurrentDictionary - The new ConcurrentDictionary from System.Collections.Concurrent that is supposed to be optimized to allow multiple threads to access safely. So the approach to each of these is also fairly straight-forward.  Let's look at the GetDelegate and AddDelegate implementations for the Dictionary with mutex lock: 1: var addDelegate = (key,val) => 2: { 3: lock (_mutex) 4: { 5: _dictionary[key] = val; 6: } 7: }; 8: var getDelegate = (key) => 9: { 10: lock (_mutex) 11: { 12: string val; 13: return _dictionary.TryGetValue(key, out val) ? val : null; 14: } 15: }; Nothing new or fancy here, just your basic lock on a private object and then query/insert into the Dictionary. Now, for the Dictionary with ReadWriteLockSlim it's a little more complex: 1: var addDelegate = (key,val) => 2: { 3: _readerWriterLock.EnterWriteLock(); 4: _dictionary[key] = val; 5: _readerWriterLock.ExitWriteLock(); 6: }; 7: var getDelegate = (key) => 8: { 9: string val; 10: _readerWriterLock.EnterReadLock(); 11: if(!_dictionary.TryGetValue(key, out val)) 12: { 13: val = null; 14: } 15: _readerWriterLock.ExitReadLock(); 16: return val; 17: }; And finally, the ConcurrentDictionary, which since it does all it's own concurrency control, is remarkably elegant and simple: 1: var addDelegate = (key,val) => 2: { 3: _concurrentDictionary[key] = val; 4: }; 5: var getDelegate = (key) => 6: { 7: string s; 8: return _concurrentDictionary.TryGetValue(key, out s) ? s : null; 9: };                    Then, I set up a test harness that would simply ask the user for the number of concurrent Accessors to attempt to Access the cache (as specified in Accessor.Access() above) and then let them fly and see how long it took them all to complete.  Each of these tests was run with 10,000,000 cache accesses divided among the available Accessor instances.  All times are in milliseconds. 1: Dictionary with Mutex Locking 2: --------------------------------------------------- 3: Accessors Mostly Misses Mostly Hits 4: 1 7916 3285 5: 10 8293 3481 6: 100 8799 3532 7: 1000 8815 3584 8:  9:  10: Dictionary with ReaderWriterLockSlim Locking 11: --------------------------------------------------- 12: Accessors Mostly Misses Mostly Hits 13: 1 8445 3624 14: 10 11002 4119 15: 100 11076 3992 16: 1000 14794 4861 17:  18:  19: Concurrent Dictionary 20: --------------------------------------------------- 21: Accessors Mostly Misses Mostly Hits 22: 1 17443 3726 23: 10 14181 1897 24: 100 15141 1994 25: 1000 17209 2128 The first test I did across the board is the Mostly Misses category.  The mostly misses (more adds because data requested was not in the dictionary) shows an interesting trend.  In both cases the Dictionary with the simple mutex lock is much faster, and the ConcurrentDictionary is the slowest solution.  But this got me thinking, and a little research seemed to confirm it, maybe the ConcurrentDictionary is more optimized to concurrent "gets" than "adds".  So since the ratio of misses to hits were 2 to 1, I decided to reverse that and see the results. So I tweaked the data so that the number of keys were much smaller than the number of iterations to give me about a 2 to 1 ration of hits to misses (twice as likely to already find the item in the cache than to need to add it).  And yes, indeed here we see that the ConcurrentDictionary is indeed faster than the standard Dictionary here.  I have a strong feeling that as the ration of hits-to-misses gets higher and higher these number gets even better as well.  This makes sense since the ConcurrentDictionary is read-optimized. Also note that I tried the tests with capacity and concurrency hints on the ConcurrentDictionary but saw very little improvement, I think this is largely because on the 10,000,000 hit test it quickly ramped up to the correct capacity and concurrency and thus the impact was limited to the first few milliseconds of the run. So what does this tell us?  Well, as in all things, ConcurrentDictionary is not a panacea.  It won't solve all your woes and it shouldn't be the only Dictionary you ever use.  So when should we use each? Use System.Collections.Generic.Dictionary when: You need a single-threaded Dictionary (no locking needed). You need a multi-threaded Dictionary that is loaded only once at creation and never modified (no locking needed). You need a multi-threaded Dictionary to store items where writes are far more prevalent than reads (locking needed). And use System.Collections.Concurrent.ConcurrentDictionary when: You need a multi-threaded Dictionary where the writes are far more prevalent than reads. You need to be able to iterate over the collection without locking it even if its being modified. Both Dictionaries have their strong suits, I have a feeling this is just one where you need to know from design what you hope to use it for and make your decision based on that criteria.

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• Robotic Arm &ndash; Hardware

  - by Szymon Kobalczyk

  This is first in series of articles about project I've been building  in my spare time since last Summer. Actually it all began when I was researching a topic of modeling human motion kinematics in order to create gesture recognition library for Kinect. This ties heavily into motion theory of robotic manipulators so I also glanced at some designs of robotic arms. Somehow I stumbled upon this cool looking open source robotic arm: It was featured on Thingiverse and published by user jjshortcut (Jan-Jaap). Since for some time I got hooked on toying with microcontrollers, robots and other electronics, I decided to give it a try and build it myself. In this post I will describe the hardware build of the arm and in later posts I will be writing about the software to control it. Another reason to build the arm myself was the cost factor. Even small commercial robotic arms are quite expensive – products from Lynxmotion and Dagu look great but both cost around USD $300 (actually there is one cheap arm available but it looks more like a toy to me). In comparison this design is quite cheap. It uses seven hobby grade servos and even the cheapest ones should work fine. The structure is build from a set of laser cut parts connected with few metal spacers (15mm and 47mm) and lots of M3 screws. Other than that you’d only need a microcontroller board to drive the servos. So in total it comes a lot cheaper to build it yourself than buy an of the shelf robotic arm. Oh, and if you don’t like this one there are few more robotic arm projects at Thingiverse (including one by oomlout). Laser cut parts Some time ago I’ve build another robot using laser cut parts so I knew the process already. You can grab the design files in both DXF and EPS format from Thingiverse, and there are also 3D models of each part in STL. Actually the design is split into a second project for the mini servo gripper (there is also a standard servo version available but it won’t fit this arm).  I wanted to make some small adjustments, layout, and add measurements to the parts before sending it for cutting. I’ve looked at some free 2D CAD programs, and finally did all this work using QCad 3 Beta with worked great for me (I also tried LibreCAD but it didn’t work that well). All parts are cut from 4 mm thick material. Because I was worried that acrylic is too fragile and might break, I also ordered another set cut from plywood. In the end I build it from plywood because it was easier to glue (I was told acrylic requires a special glue). Btw. I found a great laser cutter service in Kraków and highly recommend it (www.ebbox.com.pl). It cost me only USD $26 for both sets ($16 acrylic + $10 plywood). Metal parts I bought all the M3 screws and nuts at local hardware store. Make sure to look for nylon lock (nyloc) nuts for the gripper because otherwise it unscrews and comes apart quickly. I couldn’t find local store with metal spacers and had to order them online (you’d need 11 x 47mm and 3 x 15mm). I think I paid less than USD $10 for all metal parts. Servos This arm uses five standards size servos to drive the arm itself, and two micro servos are used on the gripper. Author of the project used Modelcraft RS-2 Servo and Modelcraft ES-05 HT Servo. I had two Futaba S3001 servos laying around, and ordered additional TowerPro SG-5010 standard size servos and TowerPro SG90 micro servos. However it turned out that the SG90 won’t fit in the gripper so I had to replace it with a slightly smaller E-Sky EK2-0508 micro servo. Later it also turned out that Futaba servos make some strange noise while working so I swapped one with TowerPro SG-5010 which has higher torque (8kg / cm). I’ve also bought three servo extension cables. All servos cost me USD $45. Assembly The build process is not difficult but you need to think carefully about order of assembling it. You can do the base and upper arm first. Because two servos in the base are close together you need to put first with one piece of lower arm already connected before you put the second servo. Then you connect the upper arm and finally put the second piece of lower arm to hold it together. Gripper and base require some gluing so think it through too. Make sure to look closely at all the photos on Thingiverse (also other people copies) and read additional posts on jjshortcust’s blog: My mini servo grippers and completed robotic arm  Multiply the robotic arm and electronics Here is also Rob’s copy cut from aluminum My assembled arm looks like this – I think it turned out really nice: Servo controller board The last piece of hardware I needed was an electronic board that would take command from PC and drive all seven servos. I could probably use Arduino for this task, and in fact there are several Arduino servo shields available (for example from Adafruit or Renbotics).  However one problem is that most support only up to six servos, and second that their accuracy is limited by Arduino’s timer frequency. So instead I looked for dedicated servo controller and found a series of Maestro boards from Pololu. I picked the Pololu Mini Maestro 12-Channel USB Servo Controller. It has many nice features including native USB connection, high resolution pulses (0.25µs) with no jitter, built-in speed and acceleration control, and even scripting capability. Another cool feature is that besides servo control, each channel can be configured as either general input or output. So far I’m using seven channels so I still have five available to connect some sensors (for example distance sensor mounted on gripper might be useful). And last but important factor was that they have SDK in .NET – what more I could wish for! The board itself is very small – half of the size of Tic-Tac box. I picked one for about USD $35 in this store. Perhaps another good alternative would be the Phidgets Advanced Servo 8-Motor – but it is significantly more expensive at USD $87.30. The Maestro Controller Driver and Software package includes Maestro Control Center program with lets you immediately configure the board. For each servo I first figured out their move range and set the min/max limits. I played with setting the speed an acceleration values as well. Big issue for me was that there are two servos that control position of lower arm (shoulder joint), and both have to be moved at the same time. This is where the scripting feature of Pololu board turned out very helpful. I wrote a script that synchronizes position of second servo with first one – so now I only need to move one servo and other will follow automatically. This turned out tricky because I couldn’t find simple offset mapping of the move range for each servo – I had to divide it into several sub-ranges and map each individually. The scripting language is bit assembler-like but gets the job done. And there is even a runtime debugging and stack view available. Altogether I’m very happy with the Pololu Mini Maestro Servo Controller, and with this final piece I completed the build and was able to move my arm from the Meastro Control program.   The total cost of my robotic arm was: $10 laser cut parts $10 metal parts $45 servos $35 servo controller ----------------------- $100 total So here you have all the information about the hardware. In next post I’ll start talking about the software that I wrote in Microsoft Robotics Developer Studio 4. Stay tuned!

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• A pseudo-listener for AlwaysOn Availability Groups for SQL Server virtual machines running in Azure

  - by MikeD

  I am involved in a project that is implementing SharePoint 2013 on virtual machines hosted in Azure. The back end data tier consists of two Azure VMs running SQL Server 2012, with the SharePoint databases contained in an AlwaysOn Availability Group. I used this "Tutorial: AlwaysOn Availability Groups in Windows Azure (GUI)" to help me implement this setup.Because Azure DHCP will not assign multiple unique IP addresses to the same VM, having an AG Listener in Azure is not currently supported.  I wanted to figure out another mechanism to support a "pseudo listener" of some sort. First, I created a CNAME (alias) record in the DNS zone with a short TTL (time to live) of 5 minutes (I may yet make this even shorter). The record represents a logical name (let's say the alias is SPSQL) of the server to connect to for the databases in the availability group (AG). When Server1 was hosting the primary replica of the AG, I would set the CNAME of SPSQL to be SERVER1. When the AG failed over to Server1, I wanted to set the CNAME to SERVER2. Seemed simple enough.(It's important to point out that the connection strings for my SharePoint services should use the CNAME alias, and not the actual server name. This whole thing falls apart otherwise.)To accomplish this, I created identical SQL Agent Jobs on Server1 and Server2, with two steps:1. Step 1: Determine if this server is hosting the primary replica.This is a TSQL step using this script:declare @agName sysname = 'AGTest'set nocount on declare @primaryReplica sysnameselect @primaryReplica = agState.primary_replicafrom sys.dm_hadr_availability_group_states agState   join sys.availability_groups ag on agstate.group_id = ag.group_id   where ag.name = @AGname if not exists(   select *    from sys.dm_hadr_availability_group_states agState   join sys.availability_groups ag on agstate.group_id = ag.group_id   where @@Servername = agstate.primary_replica    and ag.name = @AGname)begin   raiserror ('Primary replica of %s is not hosted on %s, it is hosted on %s',17,1,@Agname, @@Servername, @primaryReplica) endThis script determines if the primary replica value of the AG group is the same as the server name, which means that our server is hosting the current AG (you should update the value of the @AgName variable to the name of your AG). If this is true, I want the DNS alias to point to this server. If the current server is not hosting the primary replica, then the script raises an error. Also, if the script can't be executed because it cannot connect to the server, that also will generate an error. For the job step settings, I set the On Failure option to "Quit the job reporting success". The next step in the job will set the DNS alias to this server name, and I only want to do that if I know that it is the current primary replica, otherwise I don't want to do anything. I also include the step output in the job history so I can see the error message.Job Step 2: Update the CNAME entry in DNS with this server's name.I used a PowerShell script to accomplish this:$cname = "SPSQL.contoso.com"$query = "Select * from MicrosoftDNS_CNAMEType"$dns1 = "dc01.contoso.com"$dns2 = "dc02.contoso.com"if ((Test-Connection -ComputerName $dns1 -Count 1 -Quiet) -eq $true){    $dnsServer = $dns1}elseif ((Test-Connection -ComputerName $dns2 -Count 1 -Quiet) -eq $true) {   $dnsServer = $dns2}else{  $msg = "Unable to connect to DNS servers: " + $dns1 + ", " + $dns2   Throw $msg}$record = Get-WmiObject -Namespace "root\microsoftdns" -Query $query -ComputerName $dnsServer  | ? { $_.Ownername -match $cname }$thisServer = [System.Net.Dns]::GetHostEntry("LocalHost").HostName + "."$currentServer = $record.RecordData if ($currentServer -eq $thisServer ) {     $cname + " CNAME is up to date: " + $currentServer}else{    $cname + " CNAME is being updated to " + $thisServer + ". It was " + $currentServer    $record.RecordData = $thisServer    $record.put()}This script does a few things:finds a responsive domain controller (Test-Connection does a ping and returns a Boolean value if you specify the -Quiet parameter)makes a WMI call to the domain controller to get the current CNAME record value (Get-WmiObject)gets the FQDN of this server (GetHostEntry)checks if the CNAME record is correct and updates it if necessary(You should update the values of the variables $cname, $dns1 and $dns2 for your environment.)Since my domain controllers are also hosted in Azure VMs, either one of them could be down at any point in time, so I need to find a DC that is responsive before attempting the DNS call. The other little thing here is that the CNAME record contains the FQDN of a machine, plus it ends with a period. So the comparison of the CNAME record has to take the trailing period into account. When I tested this step, I was getting ACCESS DENIED responses from PowerShell for the Get-WmiObject cmdlet that does a remote lookup on the DC. This occurred because the SQL Agent service account was not a member of the Domain Admins group, so I decided to create a SQL Credential to store the credentials for a domain administrator account and use it as a PowerShell proxy (rather than give the service account Domain Admins membership).In SQL Management Studio, right click on the Credentials node (under the server's Security node), and choose New Credential...Then, under SQL Agent-->Proxies, right click on the PowerShell node and choose New Proxy...Finally, in the job step properties for the PowerShell step, select the new proxy in the Run As drop down.I created this two step Job on both nodes of the Availability Group, but if you had more than two nodes, just create the same job on all the servers. I set the schedule for the job to execute every minute.When the server that is hosting the primary replica is running the job, the job history looks like this:The job history on the secondary server looks like this: When a failover occurs, the SQL Agent job on the new primary replica will detect that the CNAME needs to be updated within a minute. Based on the TTL of the CNAME (which I said at the beginning was 5 minutes), the SharePoint servers will get the new alias within five minutes and should be able to reconnect. I may want to shorten up the TTL to reduce the time it takes for the client connections to use the new alias. Using a DNS CNAME and a SQL Agent Job on all servers hosting AG replicas, I was able to create a pseudo-listener to automatically change the name of the server that was hosting the primary replica, for a scenario where I cannot use a regular AG listener (in this case, because the servers are all hosted in Azure).    

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• How can a code editor effectively hint at code nesting level - without using indentation?

  - by pgfearo

  I've written an XML text editor that provides 2 view options for the same XML text, one indented (virtually), the other left-justified. The motivation for the left-justified view is to help users 'see' the whitespace characters they're using for indentation of plain-text or XPath code without interference from indentation that is an automated side-effect of the XML context. I want to provide visual clues (in the non-editable part of the editor) for the left-justified mode that will help the user, but without getting too elaborate. I tried just using connecting lines, but that seemed too busy. The best I've come up with so far is shown in a mocked up screenshot of the editor below, but I'm seeking better/simpler alternatives (that don't require too much code). [Edit] Taking the heatmap idea (from: @jimp) I get this and 3 alternatives - labelled a, b and c: The following section describes the accepted answer as a proposal, bringing together ideas from a number of other answers and comments. As this question is now community wiki, please feel free to update this. NestView The name for this idea which provides a visual method to improve the readability of nested code without using indentation. Contour Lines The name for the differently shaded lines within the NestView The image above shows the NestView used to help visualise an XML snippet. Though XML is used for this illustration, any other code syntax that uses nesting could have been used for this illustration. An Overview: The contour lines are shaded (as in a heatmap) to convey nesting level The contour lines are angled to show when a nesting level is being either opened or closed. A contour line links the start of a nesting level to the corresponding end. The combined width of contour lines give a visual impression of nesting level, in addition to the heatmap. The width of the NestView may be manually resizable, but should not change as the code changes. Contour lines can either be compressed or truncated to keep acheive this. Blank lines are sometimes used code to break up text into more digestable chunks. Such lines could trigger special behaviour in the NestView. For example the heatmap could be reset or a background color contour line used, or both. One or more contour lines associated with the currently selected code can be highlighted. The contour line associated with the selected code level would be emphasized the most, but other contour lines could also 'light up' in addition to help highlight the containing nested group Different behaviors (such as code folding or code selection) can be associated with clicking/double-clicking on a Contour Line. Different parts of a contour line (leading, middle or trailing edge) may have different dynamic behaviors associated. Tooltips can be shown on a mouse hover event over a contour line The NestView is updated continously as the code is edited. Where nesting is not well-balanced assumptions can be made where the nesting level should end, but the associated temporary contour lines must be highlighted in some way as a warning. Drag and drop behaviors of Contour Lines can be supported. Behaviour may vary according to the part of the contour line being dragged. Features commonly found in the left margin such as line numbering and colour highlighting for errors and change state could overlay the NestView. Additional Functionality The proposal addresses a range of additional issues - many are outside the scope of the original question, but a useful side-effect. Visually linking the start and end of a nested region The contour lines connect the start and end of each nested level Highlighting the context of the currently selected line As code is selected, the associated nest-level in the NestView can be highlighted Differentiating between code regions at the same nesting level In the case of XML different hues could be used for different namespaces. Programming languages (such as c#) support named regions that could be used in a similar way. Dividing areas within a nesting area into different visual blocks Extra lines are often inserted into code to aid readability. Such empty lines could be used to reset the saturation level of the NestView's contour lines. Multi-Column Code View Code without indentation makes the use of a multi-column view more effective because word-wrap or horizontal scrolling is less likely to be required. In this view, once code has reach the bottom of one column, it flows into the next one: Usage beyond merely providing a visual aid As proposed in the overview, the NestView could provide a range of editing and selection features which would be broadly in line with what is expected from a TreeView control. The key difference is that a typical TreeView node has 2 parts: an expander and the node icon. A NestView contour line can have as many as 3 parts: an opener (sloping), a connector (vertical) and a close (sloping). On Indentation The NestView presented alongside non-indented code complements, but is unlikely to replace, the conventional indented code view. It's likely that any solutions adopting a NestView, will provide a method to switch seamlessly between indented and non-indented code views without affecting any of the code text itself - including whitespace characters. One technique for the indented view would be 'Virtual Formatting' - where a dynamic left-margin is used in lieu of tab or space characters. The same nesting-level data used to dynamically render the NestView could also used for the more conventional-looking indented view. Printing Indentation will be important for the readability of printed code. Here, the absence of tab/space characters and a dynamic left-margin means that the text can wrap at the right-margin and still maintain the integrity of the indented view. Line numbers can be used as visual markers that indicate where code is word-wrapped and also the exact position of indentation: Screen Real-Estate: Flat Vs Indented Addressing the question of whether the NestView uses up valuable screen real-estate: Contour lines work well with a width the same as the code editor's character width. A NestView width of 12 character widths can therefore accommodate 12 levels of nesting before contour lines are truncated/compressed. If an indented view uses 3 character-widths for each nesting level then space is saved until nesting reaches 4 levels of nesting, after this nesting level the flat view has a space-saving advantage that increases with each nesting level. Note: A minimum indentation of 4 character widths is often recommended for code, however XML often manages with less. Also, Virtual Formatting permits less indentation to be used because there's no risk of alignment issues A comparison of the 2 views is shown below: Based on the above, its probably fair to conclude that view style choice will be based on factors other than screen real-estate. The one exception is where screen space is at a premium, for example on a Netbook/Tablet or when multiple code windows are open. In these cases, the resizable NestView would seem to be a clear winner. Use Cases Examples of real-world examples where NestView may be a useful option: Where screen real-estate is at a premium a. On devices such as tablets, notepads and smartphones b. When showing code on websites c. When multiple code windows need to be visible on the desktop simultaneously Where consistent whitespace indentation of text within code is a priority For reviewing deeply nested code. For example where sub-languages (e.g. Linq in C# or XPath in XSLT) might cause high levels of nesting. Accessibility Resizing and color options must be provided to aid those with visual impairments, and also to suit environmental conditions and personal preferences: Compatability of edited code with other systems A solution incorporating a NestView option should ideally be capable of stripping leading tab and space characters (identified as only having a formatting role) from imported code. Then, once stripped, the code could be rendered neatly in both the left-justified and indented views without change. For many users relying on systems such as merging and diff tools that are not whitespace-aware this will be a major concern (if not a complete show-stopper). Other Works: Visualisation of Overlapping Markup Published research by Wendell Piez, dated from 2004, addresses the issue of the visualisation of overlapping markup, specifically LMNL. This includes SVG graphics with significant similarities to the NestView proposal, as such, they are acknowledged here. The visual differences are clear in the images (below), the key functional distinction is that NestView is intended only for well-nested XML or code, whereas Wendell Piez's graphics are designed to represent overlapped nesting. The graphics above were reproduced - with kind permission - from http://www.piez.org Sources: Towards Hermenutic Markup Half-steps toward LMNL

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