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  • a non recursive approach to the problem of generating combinations at fault

    - by mark
    Hi, I wanted a non recursive approach to the problem of generating combination of certain set of characters or numbers. So, given a subset k of numbers n, generate all the possible combination n!/k!(n-k)! The recursive method would give a combination, given the previous one combination. A non recursive method would generate a combination of a given value of loop index i. I approached the problem with this code: Tested with n = 4 and k = 3, and it works, but if I change k to a number 3 it does not work. Is it due to the fact that (n-k)! in case of n = 4 and k = 3 is 1. and if k 3 it will be more than 1? Thanks. int facto(int x); int len,fact,rem=0,pos=0; int str[7]; int avail[7]; str[0] = 1; str[1] = 2; str[2] = 3; str[3] = 4; str[4] = 5; str[5] = 6; str[6] = 7; int tot=facto(n) / facto(n-k) / facto(k); for (int i=0;i<tot;i++) { avail[0]=1; avail[1]=2; avail[2]=3; avail[3]=4; avail[4]=5; avail[5]=6; avail[6]=7; rem = facto(i+1)-1; cout<<rem+1<<". "; for(int j=len;j>0;j--) { int div = facto(j); pos = rem / div; rem = rem % div; cout<<avail[pos]<<" "; avail[pos]=avail[j]; } cout<<endl; } int facto(int x) { int fact=1; while(x0) fact*=x--; return fact; }

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  • Lawler's Algorithm Implementation Assistance

    - by Richard Knop
    Here is my implemenation of Lawler's algorithm in PHP (I know... but I'm used to it): <?php $jobs = array(1, 2, 3, 4, 5, 6); $jobsSubset = array(2, 5, 6); $n = count($jobs); $processingTimes = array(2, 3, 4, 3, 2, 1); $dueDates = array(3, 15, 9, 7, 11, 20); $optimalSchedule = array(); foreach ($jobs as $j) { $optimalSchedule[] = 0; } $dicreasedCardinality = array(); for ($i = $n; $i >= 1; $i--) { $x = 0; $max = 0; // loop through all jobs for ($j = 0; $j < $i; $j++) { // ignore if $j already is in the $dicreasedCardinality array if (false === in_array($j, $dicreasedCardinality)) { // if the job has no succesor in $jobsSubset if (false === isset($jobs[$j+1]) || false === in_array($jobs[$j+1], $jobsSubset)) { // here I find an array index of a job with the maximum due date // amongst jobs with no sucessor in $jobsSubset if ($x < $dueDates[$j]) { $x = $dueDates[$j]; $max = $j; } } } } // move the job at the end of $optimalSchedule $optimalSchedule[$i-1] = $jobs[$max]; // decrease the cardinality of $jobs $dicreasedCardinality[] = $max; } print_r($optimalSchedule); Now the above returns an optimal schedule like this: Array ( [0] => 1 [1] => 1 [2] => 1 [3] => 3 [4] => 2 [5] => 6 ) Which doesn't seem right to me. The problem might be with my implementation of the algorithm because I am not sure I understand it correctly. I used this source to implement it: http://www.google.com/books?id=aSiBs6PDm9AC&pg=PA166&dq=lawler%27s+algorithm+code&lr=&hl=sk&cd=4#v=onepage&q=&f=false The description there is a little confusing. For example, I didn't quite get how is the subset D defined (I guess it is arbitrary). Could anyone help me out with this? I have been trying to find some sources with simpler explanation of the algorithm but all sources I found were even more complicated (with math proofs and such) so I am stuck with the link above. Yes, this is a homework, if it wasn't obvious. I still have few weeks to crack this but I have spent few days already trying to get how exactly this algorithm works with no success so I don't think I will get any brighter during that time.

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  • Corner Cases, Unexpected and Unusual Matlab

    - by Mikhail
    Over the years, reading others code, I encountered and collected some examples of Matlab syntax which can be at first unusual and counterintuitive. Please, feel free to comment or complement this list. I verified it r2006a. set([], 'Background:Color','red') Matlab is very forgiving sometimes. In this case, setting properties to an array of objects works also with nonsense properties, at least when the array is empty. myArray([1,round(end/2)]) This use of end keyword may seem unclean but is sometimes very handy instead of using length(myArray). any([]) ~= all([]) Surprisigly any([]) returns false and all([]) returns true. And I always thought that all is stronger then any. EDIT: with not empty argument all() returns true for a subset of values for which any() returns true (e.g. truth table). This means that any() false implies all() false. This simple rule is being violated by Matlab with [] as argument. Loren also blogged about it. Select(Range(ExcelComObj)) Procedural style COM object method dispatch. Do not wonder that exist('Select') returns zero! [myString, myCell] Matlab makes in this case an implicit cast of string variable myString to cell type {myString}. It works, also if I would not expect it to do so. [double(1.8), uint8(123)] => 2 123 Another cast example. Everybody would probably expect uint8 value being cast to double but Mathworks have another opinion. a = 5; b = a(); It looks silly but you can call a variable with round brackets. Actually it makes sense because this way you can execute a function given its handle. a = {'aa', 'bb' 'cc', 'dd'}; Surprsisingly this code neither returns a vector nor rises an error but defins matrix, using just code layout. It is probably a relict from ancient times. set(hobj, {'BackgroundColor','ForegroundColor'},{'red','blue'}) This code does what you probably expect it to do. That function set accepts a struct as its second argument is a known fact and makes sense, and this sintax is just a cell2struct away. Equvalence rules are sometimes unexpected at first. For example 'A'==65 returns true (although for C-experts it is self-evident). About which further unexpected/unusual Matlab features are you aware?

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  • How can I inject Javascript (including Prototype.js) in other sites without cluttering the global na

    - by Daniel Magliola
    I'm currently on a project that is a big site that uses the Prototype library, and there is already a humongous amount of Javascript code. We're now working on a piece of code that will get "injected" into other people's sites (picture people adding a <script> tag in their sites) which will then run our code and add a bunch of DOM elements and functionality to their site. This will have new pieces of code, and will also reuse a lot of the code that we use on our main site. The problem I have is that it's of course not cool to just add a <script> that will include Prototype in people's pages. If we do that in a page that's already using ANY framework, we're guaranteed to screw everything up. jQuery gives us the option to "rename" the $ object, so it could handle this situation decently, except obviously for the fact that we're not using jQuery, so we'd have to migrate everything. Right now i'm contemplating a number of ugly choices, and I'm not sure what's best... Rewrite everything to use jQuery, with a renamed $ object everywhere. Creating a "new" Prototype library with only the subset we'd be using in "injected" code, and renaming $ to something else. Then again I'd have to adapt the parts of my code that would be shared somehow. Not using a library at all in injected code, to keep it as clean as possible, and rewriting the shared code to use no library at all. This would obviously degenerate into us creating our own frankenstein of a library, which is probably the worst case scenario ever. I'm wondering what you guys think I could do, and also whether there's some magic option that would solve all my problems... For example, do you think I could use something like Caja / Cajita to sandbox my own code and isolate it from the rest of the site, and have Prototype inside of there? Or am I completely missing the point with that? I also read once about a technique for bookmarklets, were you add your code like this: (function() { /* your code */ })(); And then your code is all inside your anonymous function and you haven't touched the global namespace at all. Do you think I could make one file containing: (function() { /* Full Code of the Prototype file here */ /* All my code that will run in the "other" site */ InitializeStuff_CreateDOMElements_AttachEventHandlers(); })(); Would that work? Would it accomplish the objective of not cluttering the global namespace, and not killing the functionality on a site that uses jQuery, for example? Or is Prototype too complex somehow to isolate it like that? (NOTE: I think I know that that would create closures everywhere and that's slower, but I don't care too much about performance, my code is not doing anything that complex)

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  • What's the best way to do base36 arithmetic in perl?

    - by DVK
    What's the best way to do base36 arithmetic in Perl? To be more specific, I need to be able to do the following: Operate on positive N-digit numbers in base 36 (e.g. digits are 0-9 A-Z) N is finite, say 9 Provide basic arithmetic, at the very least the following 3: Addition (A+B) Subtraction (A-B) Whole division, e.g. floor(A/B). Strictly speaking, I don't really need a base10 conversion ability - the numbers will 100% of time be in base36. So I'm quite OK if the solution does NOT implement conversion from base36 back to base10 and vice versa. I don't much care whether the solution is brute-force "convert to base 10 and back" or converting to binary, or some more elegant approach "natively" performing baseN operations (as stated above, to/from base10 conversion is not a requirement). My only 3 considerations are: It fits the minimum specifications above It's "standard". Currently we're using and old homegrown module based on base10 conversion done by hand that is buggy and sucks. I'd much rather replace that with some commonly used CPAN solution instead of re-writing my own bicycle from scratch, but I'm perfectly capable of building it if no better standard possibility exists. It must be fast-ish (though not lightning fast). Something that takes 1 second to sum up 2 9-digit base36 numbers is worse than anything I can roll on my own :) P.S. Just to provide some context in case people decide to solve my XY problem for me in addition to answering the technical question above :) We have a fairly large tree (stored in DB as a bunch of edges), and we need to superimpose order on a subset of that tree. The tree dimentions are big both depth- and breadth- wise. The tree is VERY actively updated (inserts and deletes and branch moves). This is currently done by having a second table with 3 columns: parent_vertex, child_vertex, local_order, where local_order is an 9-character string built of A-Z0-9 (e.g. base 36 number). Additional considerations: It is required that the local order is unique per child (and obviously unique per parent), Any complete re-ordering of a parent is somewhat expensive, and thus the implementation is to try and assign - for a parent with X children - the orders which are somewhat evenly distributed between 0 and 36**10-1, so that almost no tree inserts result in a full re-ordering.

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  • Mapping two tables 0..n in Hibernate

    - by simon
    I have a table Users CREATE TABLE "USERS" ( "ID" NUMBER NOT NULL , "LOGINNAME" VARCHAR2 (150) NOT NULL ) and I have a second table SpecialUsers. No UserId can occur twice in the SpecialUsers table, and only a small subset of the ids of users in the Users table are contained in the SpecialUsers table. CREATE TABLE "SPECIALUSERS" ( "USERID" NUMBER NOT NULL, CONSTRAINT "PK_SPECIALUSERS" PRIMARY KEY ("USERID") ) ALTER TABLE "SPECIALUSERS" ADD CONSTRAINT "FK_SPECIALUSERS_USERID" FOREIGN KEY ("USERID") REFERENCES "USERS" ("ID") / Mapping the Users table in Hibernate works ok <hibernate-mapping package="com.initech.domain"> <class name="com.initech.User" table="USERS"> <id name="id" column="ID" type="java.lang.Long"> <meta attribute="use-in-tostring">true</meta> <generator class="sequence"> <param name="sequence">SEQ_USERS_ID</param> </generator> </id> <property name="loginName" column="LOGINNAME" type="java.lang.String" not-null="true"> <meta attribute="use-in-tostring">true</meta> </property> </class> </hibernate-mapping> But I'm struggling in creating the mapping for the SpecialUsers table. All the examples (e.g. in Hibernate documentation) in Internet I found don't have this type of self-reference. I tried a mapping like this: <hibernate-mapping package="com.initech.domain"> <class name="com.initech.User" table="SPECIALUSERS"> <id name="id" column="USERID"> <meta attribute="use-in-tostring">true</meta> <generator class="foreign"> <param name="property">user</param> </generator> </id> <one-to-one name="user" class="User"/> </class> </hibernate-mapping> but got the error Invocation of init method failed; nested exception is org.hibernate.DuplicateMappingException: Duplicate class/entity mapping com.initech.User How should I map the SpecialUsers table? What I need on the application level is a list of the User objects contained in the SpecialUsers table.

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  • Write Scheme data structures so they can be eval-d back in, or alternative

    - by Jesse Millikan
    I'm writing an application (A juggling pattern animator) in PLT Scheme that accepts Scheme expressions as values for some fields. I'm attempting to write a small text editor that will let me "explode" expressions into expressions that can still be eval'd but contain the data as literals for manual tweaking. For example, (4hss->sexp "747") is a function call that generates a legitimate pattern. If I eval and print that, it becomes (((7 3) - - -) (- - (4 2) -) (- (7 2) - -) (- - - (7 1)) ((4 0) - - -) (- - (7 0) -) (- (7 2) - -) (- - - (4 3)) ((7 3) - - -) (- - (7 0) -) (- (4 1) - -) (- - - (7 1))) which can be "read" as a string, but will not "eval" the same as the function. For this statement, of course, what I need would be as simple as (quote (((7 3... but other examples are non-trivial. This one, for example, contains structs which print as vectors: pair-of-jugglers ; --> (#(struct:hand #(struct:position -0.35 2.0 1.0) #(struct:position -0.6 2.05 1.1) 1.832595714594046) #(struct:hand #(struct:position 0.35 2.0 1.0) #(struct:position 0.6 2.0500000000000003 1.1) 1.308996938995747) #(struct:hand #(struct:position 0.35 -2.0 1.0) #(struct:position 0.6 -2.05 1.1) -1.3089969389957472) #(struct:hand #(struct:position -0.35 -2.0 1.0) #(struct:position -0.6 -2.05 1.1) -1.8325957145940461)) I've thought of at least three possible solutions, none of which I like very much. Solution A is to write a recursive eval-able output function myself for a reasonably large subset of the values that I might be using. There (probably...) won't be any circular references by the nature of the data structures used, so that wouldn't be such a long job. The output would end up looking like `(((3 0) (... ; ex 1 `(,(make-hand (make-position ... ; ex 2 Or even worse if I could't figure out how to do it properly with quasiquoting. Solution B would be to write out everything as (read (open-input-string "(big-long-s-expression)")) which, technically, solves the problem I'm bringing up but is... ugly. Solution C might be a different approach of giving up eval and using only read for parsing input, or an uglier approach where the s-expression is used as directly data if eval fails, but those both seem unpleasant compared to using scheme values directly. Undiscovered Solution D would be a PLT Scheme option, function or library I haven't located that would match Solution A. Help me out before I start having bad recursion dreams again.

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  • Algorithm for finding symmetries of a tree

    - by Paxinum
    I have n sectors, enumerated 0 to n-1 counterclockwise. The boundaries between these sectors are infinite branches (n of them). The sectors live in the complex plane, and for n even, sector 0 and n/2 are bisected by the real axis, and the sectors are evenly spaced. These branches meet at certain points, called junctions. Each junction is adjacent to a subset of the sectors (at least 3 of them). Specifying the junctions, (in pre-fix order, lets say, starting from junction adjacent to sector 0 and 1), and the distance between the junctions, uniquely describes the tree. Now, given such a representation, how can I see if it is symmetric wrt the real axis? For example, n=6, the tree (0,1,5)(1,2,4,5)(2,3,4) have three junctions on the real line, so it is symmetric wrt the real axis. If the distances between (015) and (1245) is equal to distance from (1245) to (234), this is also symmetric wrt the imaginary axis. The tree (0,1,5)(1,2,5)(2,4,5)(2,3,4) have 4 junctions, and this is never symmetric wrt either imaginary or real axis, but it has 180 degrees rotation symmetry if the distance between the first two and the last two junctions in the representation are equal. Edit: This is actually for my research. I have posted the question at mathoverflow as well, but my days in competition programming tells me that this is more like an IOI task. Code in mathematica would be excellent, but java, python, or any other language readable by a human suffices. Here are some examples (pretend the double edges are single and we have a tree) http://www2.math.su.se/~per/files.php?file=contr_ex_1.pdf http://www2.math.su.se/~per/files.php?file=contr_ex_2.pdf http://www2.math.su.se/~per/files.php?file=contr_ex_5.pdf Example 1 is described as (0,1,4)(1,2,4)(2,3,4)(0,4,5) with distances (2,1,3). Example 2 is described as (0,1,4)(1,2,4)(2,3,4)(0,4,5) with distances (2,1,1). Example 5 is described as (0,1,4,5)(1,2,3,4) with distances (2). So, given the description/representation, I want to find some algorithm to decide if it is symmetric wrt real, imaginary, and rotation 180 degrees. The last example have 180 degree symmetry. (These symmetries corresponds to special kinds of potential in the Schroedinger equation, which has nice properties in quantum mechanics.)

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  • How to optimize this SQL query for a rectangular region?

    - by Andrew B.
    I'm trying to optimize the following query, but it's not clear to me what index or indexes would be best. I'm storing tiles in a two-dimensional plane and querying for rectangular regions of that plane. The table has, for the purposes of this question, the following columns: id: a primary key integer world_id: an integer foreign key which acts as a namespace for a subset of tiles tileY: the Y-coordinate integer tileX: the X-coordinate integer value: the contents of this tile, a varchar if it matters. I have the following indexes: "ywot_tile_pkey" PRIMARY KEY, btree (id) "ywot_tile_world_id_key" UNIQUE, btree (world_id, "tileY", "tileX") "ywot_tile_world_id" btree (world_id) And this is the query I'm trying to optimize: ywot=> EXPLAIN ANALYZE SELECT * FROM "ywot_tile" WHERE ("world_id" = 27685 AND "tileY" <= 6 AND "tileX" <= 9 AND "tileX" >= -2 AND "tileY" >= -1 ); QUERY PLAN ------------------------------------------------------------------------------------------------------------------------------------------- Bitmap Heap Scan on ywot_tile (cost=11384.13..149421.27 rows=65989 width=168) (actual time=79.646..80.075 rows=96 loops=1) Recheck Cond: ((world_id = 27685) AND ("tileY" <= 6) AND ("tileY" >= (-1)) AND ("tileX" <= 9) AND ("tileX" >= (-2))) -> Bitmap Index Scan on ywot_tile_world_id_key (cost=0.00..11367.63 rows=65989 width=0) (actual time=79.615..79.615 rows=125 loops=1) Index Cond: ((world_id = 27685) AND ("tileY" <= 6) AND ("tileY" >= (-1)) AND ("tileX" <= 9) AND ("tileX" >= (-2))) Total runtime: 80.194 ms So the world is fixed, and we are querying for a rectangular region of tiles. Some more information that might be relevant: All the tiles for a queried region may or may not be present The height and width of a queried rectangle are typically about 10x10-20x20 For any given (world, X) or (world, Y) pair, there may be an unbounded number of matching tiles, but the worst case is currently around 10,000, and typically there are far fewer. New tiles are created far less frequently than existing ones are updated (changing the 'value'), and that itself is far less frequent that just reading as in the query above. The only thing I can think of would be to index on (world, X) and (world, Y). My guess is that the database would be able to take those two sets and intersect them. The problem is that there is a potentially unbounded number of matches for either for either of those. Is there some other kind of index that would be more appropriate?

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  • What to name column in database table that holds versioning number

    - by rwmnau
    I'm trying to figure out what to call the column in my database table that holds an INT to specific "record version". I'm currently using "RecordOrder", but I don't like that, because people think higher=newer, but the way I'm using it, lower=newer (with "1" being the current record, "2" being the second most current, "3" older still, and so on). I've considered "RecordVersion", but I'm afraid that would have the same problem. Any other suggestions? "RecordAge"? I'm doing this because when I insert into the table, instead of having to find out what version is next, then run the risk of having that number stolen from me before I write, I just insert insert with a "RecordOrder" of 0. There's a trigger on the table AFTER INSERT that increments all the "RecordOrder" numbers for that key by 1, so the record I just inserted becomes "1", and all others are increased by 1. That way, you can get a person's current record by selection RecordOrder=1, instead of getting the MAX(RecordOrder) and then selecting that. PS - I'm also open to criticism about why this is a terrible idea and I should be incrementing this index instead. This just seemed to make lookups much easier, but if it's a bad idea, please enlighten me! Some details about the data, as an example: I have the following database table: CREATE TABLE AmountDue ( CustomerNumber INT, AmountDue DECIMAL(14,2), RecordOrder SMALLINT, RecordCreated DATETIME ) A subset of my data looks like this: CustomerNumber Amountdue RecordOrder RecordCreated 100 0 1 2009-12-19 05:10:10.123 100 10.05 2 2009-12-15 06:12:10.123 100 100.00 3 2009-12-14 14:19:10.123 101 5.00 1 2009-11-14 05:16:10.123 In this example, there are three rows for customer 100 - they owed $100, then $10.05, and now they owe nothing. Let me know if I need to clarify it some more. UPDATE: The "RecordOrder" and "RecordCreated" columns are not available to the user - they're only there for internal use, and to help figure out which is the current customer record. Also, I could use it to return an appropriately-ordered customer history, though I could just as easily do that with the date. I can accomplish the same thing as an incrementing "Record Version" with just the RecordCreated date, I suppose, but that removes the convenience of knowing that RecordOrder=1 is the current record, and I'm back to doing a sub-query with MAX or MIN on the DateTime to determine the most recent record.

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  • If we don't like it for the presentation layer, then why do we tolerate it for the behavior layer?

    - by greim
    Suppose CSS as we know it had never been invented, and the closest we could get was to do this: <script> // this is the page's stylesheet $(document).ready(function(){ $('.error').css({'color':'red'}); $('a[href]').css({'textDecoration':'none'}); ... }); </script> If this was how we were forced to write code, would we put up with it? Or would every developer on Earth scream at browser vendors until they standardized upon CSS, or at least some kind of declarative style language? Maybe CSS isn't perfect, but hopefully it's obvious how it's better than the find things, do stuff method shown above. So my question is this. We've seen and tasted of the glory of declarative binding with CSS, so why, when it comes to the behavioral/interactive layer, does the entire JavaScript community seem complacent about continuing to use the kludgy procedural method described above? Why for example is this considered by many to be the best possible way to do things: <script> $(document).ready(function(){ $('.widget').append("<a class='button' href='#'>...</div>"); $('a[href]').click(function(){...}); ... }); </script> Why isn't there a massive push to get XBL2.0 or .htc files or some kind of declarative behavior syntax implemented in a standard way across browsers? Is this recognized as a need by other web development professionals? Is there anything on the horizon for HTML5? (Caveats, disclaimers, etc: I realize that it's not a perfect world and that we're playing the hand we've been dealt. My point isn't to criticize the current way of doing things so much as to criticize the complacency that exists about the current way of doing things. Secondly, event delegation, especially at the root level, is a step closer to having a declarative behavior layer. It solves a subset of the problem, but it can't create UI elements, so the overall problem remains.)

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  • How to learn proper C++?

    - by Chris
    While reading a long series of really, really interesting threads, I've come to a realization: I don't think I really know C++. I know C, I know classes, I know inheritance, I know templates (& the STL) and I know exceptions. Not C++. To clarify, I've been writing "C++" for more than 5 years now. I know C, and I know that C and C++ share a common subset. What I've begun to realize, though, is that more times than not, I wind up treating C++ something vaguely like "C with classes," although I do practice RAII. I've never used Boost, and have only read up on TR1 and C++0x - I haven't used any of these features in practice. I don't use namespaces. I see a list of #defines, and I think - "Gracious, that's horrible! Very un-C++-like," only to go and mindlessly write class wrappers for the sake of it, and I wind up with large numbers (maybe a few per class) of static methods, and for some reason, that just doesn't seem right lately. The professional in me yells "just get the job done," the academic yells "you should write proper C++ when writing C++" and I feel like the point of balance is somewhere in between. I'd like to note that I don't want to program "pure" C++ just for the sake of it. I know several languages. I have a good feel for what "Pythonic" is. I know what clean and clear PHP is. Good C code I can read and write better than English. The issue is that I learned C by example, and picked up C++ as a "series of modifications" to C. And a lot of my early C++ work was creating class wrappers for C libraries. I feel like my own personal C-heavy background while learning C++ has sort of... clouded my acceptance of C++ in it's own right, as it's own language. Do the weathered C++ lags here have any advice for me? Good examples of clean, sharp C++ to learn from? What habits of C does my inner-C++ really need to break from? My goal here is not to go forth and trumpet "good" C++ paradigm from rooftops for the sake of it. C and C++ are two different languages, and I want to start treating them that way. How? Where to start? Thanks in advance! Cheers, -Chris

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  • MSSQL 2005: Update rows in a specified order (like ORDER BY)?

    - by JMTyler
    I want to update rows of a table in a specific order, like one would expect if including an ORDER BY clause, but MS SQL does not support the ORDER BY clause in UPDATE queries. I have checked out this question which supplied a nice solution, but my query is a bit more complicated than the one specified there. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) ORDER BY Parent.Depth DESC; So, what I'm hoping that you'll notice is that a single table (TableA) contains a hierarchy of rows, wherein one row can be the parent or child of any other row. The rows need to be updated in order from the deepest child up to the root parent. This is because TableA.ColA must contain an up-to-date concatenation of its own current value with the values of its children (I realize this query only concats with one child, but that is for the sake of simplicity - the purpose of the example in this question does not necessitate any more verbosity), therefore the query must update from the bottom up. The solution suggested in the question I noted above is as follows: UPDATE messages SET status=10 WHERE ID in (SELECT TOP (10) Id FROM Table WHERE status=0 ORDER BY priority DESC ); The reason that I don't think I can use this solution is because I am referencing column values from the parent table inside my subquery (see WHERE Child.ParentColB = Parent.ColB), and I don't think two sibling subqueries would have access to each others' data. So far I have only determined one way to merge that suggested solution with my current problem, and I don't think it works. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) WHERE Parent.Id IN (SELECT Id FROM TableA ORDER BY Parent.Depth DESC); The WHERE..IN subquery will not actually return a subset of the rows, it will just return the full list of IDs in the order that I want. However (I don't know for sure - please tell me if I'm wrong) I think that the WHERE..IN clause will not care about the order of IDs within the parentheses - it will just check the ID of the row it currently wants to update to see if it's in that list (which, they all are) in whatever order it is already trying to update... Which would just be a total waste of cycles, because it wouldn't change anything. So, in conclusion, I have looked around and can't seem to figure out a way to update in a specified order (and included the reason I need to update in that order, because I am sure I would otherwise get the ever-so-useful "why?" answers) and I am now hitting up Stack Overflow to see if any of you gurus out there who know more about SQL than I do (which isn't saying much) know of an efficient way to do this. It's particularly important that I only use a single query to complete this action. A long question, but I wanted to cover my bases and give you guys as much info to feed off of as possible. :) Any thoughts?

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  • Converting C source to C++

    - by Barry Kelly
    How would you go about converting a reasonably large (300K), fairly mature C codebase to C++? The kind of C I have in mind is split into files roughly corresponding to modules (i.e. less granular than a typical OO class-based decomposition), using internal linkage in lieu private functions and data, and external linkage for public functions and data. Global variables are used extensively for communication between the modules. There is a very extensive integration test suite available, but no unit (i.e. module) level tests. I have in mind a general strategy: Compile everything in C++'s C subset and get that working. Convert modules into huge classes, so that all the cross-references are scoped by a class name, but leaving all functions and data as static members, and get that working. Convert huge classes into instances with appropriate constructors and initialized cross-references; replace static member accesses with indirect accesses as appropriate; and get that working. Now, approach the project as an ill-factored OO application, and write unit tests where dependencies are tractable, and decompose into separate classes where they are not; the goal here would be to move from one working program to another at each transformation. Obviously, this would be quite a bit of work. Are there any case studies / war stories out there on this kind of translation? Alternative strategies? Other useful advice? Note 1: the program is a compiler, and probably millions of other programs rely on its behaviour not changing, so wholesale rewriting is pretty much not an option. Note 2: the source is nearly 20 years old, and has perhaps 30% code churn (lines modified + added / previous total lines) per year. It is heavily maintained and extended, in other words. Thus, one of the goals would be to increase mantainability. [For the sake of the question, assume that translation into C++ is mandatory, and that leaving it in C is not an option. The point of adding this condition is to weed out the "leave it in C" answers.]

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  • What database table structure should I use for versions, codebases, deployables?

    - by Zac Thompson
    I'm having doubts about my table structure, and I wonder if there is a better approach. I've got a little database for version control repositories (e.g. SVN), the packages (e.g. Linux RPMs) built therefrom, and the versions (e.g. 1.2.3-4) thereof. A given repository might produce no packages, or several, but if there are more than one for a given repository then a particular version for that repository will indicate a single "tag" of the codebase. A particular version "string" might be used to tag a version of the source code in more than one repository, but there may be no relationship between "1.0" for two different repos. So if packages P and Q both come from repo R, then P 1.0 and Q 1.0 are both built from the 1.0 tag of repo R. But if package X comes from repo Y, then X 1.0 has no relationship to P 1.0. In my (simplified) model, I have the following tables (the x_id columns are auto-incrementing surrogate keys; you can pretend I'm using a different primary key if you wish, it's not really important): repository - repository_id - repository_name (unique) ... version - version_id - version_string (unique for a particular repository) - repository_id ... package - package_id - package_name (unique) - repository_id ... This makes it easy for me to see, for example, what are valid versions of a given package: I can join with the version table using the repository_id. However, suppose I would like to add some information to this database, e.g., to indicate which package versions have been approved for release. I certainly need a new table: package_version - version_id - package_id - package_version_released ... Again, the nature of the keys that I use are not really important to my problem, and you can imagine that the data column is "promotion_level" or something if that helps. My doubts arise when I realize that there's really a very close relationship between the version_id and the package_id in my new table ... they must share the same repository_id. Only a small subset of package/version combinations are valid. So I should have some kind of constraint on those columns, enforcing that ... ... I don't know, it just feels off, somehow. Like I'm including somehow more information than I really need? I don't know how to explain my hesitance here. I can't figure out which (if any) normal form I'm violating, but I also can't find an example of a schema with this sort of structure ... not being a DBA by profession I'm not sure where to look. So I'm asking: am I just being overly sensitive?

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  • Is a many-to-many relationship with extra fields the right tool for my job?

    - by whichhand
    Previously had a go at asking a more specific version of this question, but had trouble articulating what my question was. On reflection that made me doubt if my chosen solution was correct for the problem, so this time I will explain the problem and ask if a) I am on the right track and b) if there is a way around my current brick wall. I am currently building a web interface to enable an existing database to be interrogated by (a small number of) users. Sticking with the analogy from the docs, I have models that look something like this: class Musician(models.Model): first_name = models.CharField(max_length=50) last_name = models.CharField(max_length=50) dob = models.DateField() class Album(models.Model): artist = models.ForeignKey(Musician) name = models.CharField(max_length=100) class Instrument(models.Model): artist = models.ForeignKey(Musician) name = models.CharField(max_length=100) Where I have one central table (Musician) and several tables of associated data that are related by either ForeignKey or OneToOneFields. Users interact with the database by creating filtering criteria to select a subset of Musicians based on data the data on the main or related tables. Likewise, the users can then select what piece of data is used to rank results that are presented to them. The results are then viewed initially as a 2 dimensional table with a single row per Musician with selected data fields (or aggregates) in each column. To give you some idea of scale, the database has ~5,000 Musicians with around 20 fields of related data. Up to here is fine and I have a working implementation. However, it is important that I have the ability for a given user to upload there own annotation data sets (more than one) and then filter and order on these in the same way they can with the existing data. The way I had tried to do this was to add the models: class UserDataSets(models.Model): user = models.ForeignKey(User) name = models.CharField(max_length=100) description = models.CharField(max_length=64) results = models.ManyToManyField(Musician, through='UserData') class UserData(models.Model): artist = models.ForeignKey(Musician) dataset = models.ForeignKey(UserDataSets) score = models.IntegerField() class Meta: unique_together = (("artist", "dataset"),) I have a simple upload mechanism enabling users to upload a data set file that consists of 1 to 1 relationship between a Musician and their "score". Within a given user dataset each artist will be unique, but different datasets are independent from each other and will often contain entries for the same musician. This worked fine for displaying the data, starting from a given artist I can do something like this: artist = Musician.objects.get(pk=1) dataset = UserDataSets.objects.get(pk=5) print artist.userdata_set.get(dataset=dataset.pk) However, this approach fell over when I came to implement the filtering and ordering of query set of musicians based on the data contained in a single user data set. For example, I could easily order the query set based on all of the data in the UserData table like this: artists = Musician.objects.all().order_by(userdata__score) But that does not help me order by the results of a given single user dataset. Likewise I need to be able to filter the query set based on the "scores" from different user data sets (eg find all musicians with a score 5 in dataset1 and < 2 in dataset2). Is there a way of doing this, or am I going about the whole thing wrong?

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  • SVG via dynamic XML+XSL

    - by Daniel
    This is a bit of a vague notion which I have been running over in my head, and which I am very curious if there is an elegant method of solving. Perhaps it should be taken as a thought experiment. Imagine you have an XML schema with a corresponding XSL transform, which renders the XML as SVG in the browser. The XSL generates SVG with appropriate Javascript handlers that, ultimately, implement editing-like functionality such that properties of the objects or their locations on the SVG canvas can be edited by the user. For instance, an element can be dragged from one location to another. Now, this isn't particularly difficult - the drag/drop example is simply a matter of changing the (x,y) coordinates of the SVG object, or a resize operation would be a simple matter of changing its width or height. But is there an elegant way to have Javascript work on the DOM of the source XML document instead of the rendered SVG? Why, you ask? Well, imagine you have very complex XSL transforms, where the modification of one property results in complex changes to the SVG. You want to maintain simplicity in your Javascript code, but also a simple way to persist the modified XML back to the server. Some possibilities of how this may function: After modification of the source DOM, simply re-run the XSL transform and replace the original. Downside: brute force, potentially expensive operation. Create id/class naming conventions in the source and target XML/SVG so elements can be related back to each other, and do an XSL transform on only a subset of the new DOM. In other words, modify temporary DOM, apply XSL to it, remove changed elements from SVG, and insert the new one. Downside: May not be possible to apply XSL to temporary in-browser DOMs(?). Also, perhaps a bit convoluted or ugly to maintain. I think that it may be possible to come up with a framework that handles the second scenario, but the challenge would be making it lightweight and not heavily tied to the actual XML schema. Any ideas or other possibilities? Or is there maybe an existing method of doing this which I'm not aware of? UPDATE: To clarify, as I mentioned in a comment below, this aids in separating the draw code from the edit code. For a more concrete example of how this is useful, imagine an element which determines how it is drawn dependent on the value of a property of an adjacent element. It's better to condense that logic directly in the draw code instead of also duplicating it in the edit code.

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  • SQL Server 2005: Update rows in a specified order (like ORDER BY)?

    - by JMTyler
    I want to update rows of a table in a specific order, like one would expect if including an ORDER BY clause, but SQL Server does not support the ORDER BY clause in UPDATE queries. I have checked out this question which supplied a nice solution, but my query is a bit more complicated than the one specified there. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) ORDER BY Parent.Depth DESC; So, what I'm hoping that you'll notice is that a single table (TableA) contains a hierarchy of rows, wherein one row can be the parent or child of any other row. The rows need to be updated in order from the deepest child up to the root parent. This is because TableA.ColA must contain an up-to-date concatenation of its own current value with the values of its children (I realize this query only concats with one child, but that is for the sake of simplicity - the purpose of the example in this question does not necessitate any more verbosity), therefore the query must update from the bottom up. The solution suggested in the question I noted above is as follows: UPDATE messages SET status=10 WHERE ID in (SELECT TOP (10) Id FROM Table WHERE status=0 ORDER BY priority DESC ); The reason that I don't think I can use this solution is because I am referencing column values from the parent table inside my subquery (see WHERE Child.ParentColB = Parent.ColB), and I don't think two sibling subqueries would have access to each others' data. So far I have only determined one way to merge that suggested solution with my current problem, and I don't think it works. UPDATE TableA AS Parent SET Parent.ColA = Parent.ColA + (SELECT TOP 1 Child.ColA FROM TableA AS Child WHERE Child.ParentColB = Parent.ColB ORDER BY Child.Priority) WHERE Parent.Id IN (SELECT Id FROM TableA ORDER BY Parent.Depth DESC); The WHERE..IN subquery will not actually return a subset of the rows, it will just return the full list of IDs in the order that I want. However (I don't know for sure - please tell me if I'm wrong) I think that the WHERE..IN clause will not care about the order of IDs within the parentheses - it will just check the ID of the row it currently wants to update to see if it's in that list (which, they all are) in whatever order it is already trying to update... Which would just be a total waste of cycles, because it wouldn't change anything. So, in conclusion, I have looked around and can't seem to figure out a way to update in a specified order (and included the reason I need to update in that order, because I am sure I would otherwise get the ever-so-useful "why?" answers) and I am now hitting up Stack Overflow to see if any of you gurus out there who know more about SQL than I do (which isn't saying much) know of an efficient way to do this. It's particularly important that I only use a single query to complete this action. A long question, but I wanted to cover my bases and give you guys as much info to feed off of as possible. :) Any thoughts?

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  • R: Plotting a graph with different colors of points based on advanced criteria

    - by balconydoor
    What I would like to do is a plot (using ggplot), where the x axis represent years which have a different colour for the last three years in the plot than the rest. The last three years should also meet a certain criteria and based on this the last three years can either be red or green. The criteria is that the mean of the last three years should be less (making it green) or more (making it red) than the 66%-percentile of the remaining years. So far I have made two different functions calculating the last three year mean: LYM3 <- function (x) { LYM3 <- tail(x,3) mean(LYM3$Data,na.rm=T) } And the 66%-percentile for the remaining: perc66 <- function(x) { percentile <- head(x,-3) quantile(percentile$Data, .66, names=F,na.rm=T) } Here are two sets of data that can be used in the calculations (plots), the first which is an example from my real data where LYM3(df1) < perc66(df1) and the second is just made up data where LYM3 perc66. df1<- data.frame(Year=c(1979:2010), Data=c(347261.87, 145071.29, 110181.93, 183016.71, 210995.67, 205207.33, 103291.78, 247182.10, 152894.45, 170771.50, 206534.55, 287770.86, 223832.43, 297542.86, 267343.54, 475485.47, 224575.08, 147607.81, 171732.38, 126818.10, 165801.08, 136921.58, 136947.63, 83428.05, 144295.87, 68566.23, 59943.05, 49909.08, 52149.11, 117627.75, 132127.79, 130463.80)) df2 <- data.frame(Year=c(1979:2010), Data=c(sample(50,29,replace=T),75,75,75)) Here’s my code for my plot so far: plot <- ggplot(df1, aes(x=Year, y=Data)) + theme_bw() + geom_point(size=3, aes(colour=ifelse(df1$Year<2008, "black",ifelse(LYM3(df1) < perc66(df1),"green","red")))) + geom_line() + scale_x_continuous(breaks=c(1980,1985,1990,1995,2000,2005,2010), limits=c(1978,2011)) plot As you notice it doesn’t really do what I want it to do. The only thing it does seem to do is that it turns the years before 2008 into one level and those after into another one and base the point colour off these two levels. Since I don’t want this year to be stationary either, I made another tiny function: fun3 <- function(x) { df <- subset(x, Year==(max(Year)-2)) df$Year } So the previous code would have the same effect as: geom_point(size=3, aes(colour=ifelse(df1$Year<fun3(df1), "black","red"))) But it still does not care about my colours. Why does it make the years into levels? And how come an ifelse function doesn’t work within another one in this case? How would it be possible to the arguments to do what I like? I realise this might be a bit messy, asking for a lot at the same time, but I hope my description is pretty clear. It would be helpful if someone could at least point me in the right direction. I tried to put the code for the plot into a function as well so I wouldn’t have to change the data frame at all functions within the plot, but I can’t get it to work. Thank you!

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  • How to find same-value rectangular areas of a given size in a matrix most efficiently?

    - by neo
    My problem is very simple but I haven't found an efficient implementation yet. Suppose there is a matrix A like this: 0 0 0 0 0 0 0 4 4 2 2 2 0 0 4 4 2 2 2 0 0 0 0 2 2 2 1 1 0 0 0 0 0 1 1 Now I want to find all starting positions of rectangular areas in this matrix which have a given size. An area is a subset of A where all numbers are the same. Let's say width=2 and height=3. There are 3 areas which have this size: 2 2 2 2 0 0 2 2 2 2 0 0 2 2 2 2 0 0 The result of the function call would be a list of starting positions (x,y starting with 0) of those areas. List((2,1),(3,1),(5,0)) The following is my current implementation. "Areas" are called "surfaces" here. case class Dimension2D(width: Int, height: Int) case class Position2D(x: Int, y: Int) def findFlatSurfaces(matrix: Array[Array[Int]], surfaceSize: Dimension2D): List[Position2D] = { val matrixWidth = matrix.length val matrixHeight = matrix(0).length var resultPositions: List[Position2D] = Nil for (y <- 0 to matrixHeight - surfaceSize.height) { var x = 0 while (x <= matrixWidth - surfaceSize.width) { val topLeft = matrix(x)(y) val topRight = matrix(x + surfaceSize.width - 1)(y) val bottomLeft = matrix(x)(y + surfaceSize.height - 1) val bottomRight = matrix(x + surfaceSize.width - 1)(y + surfaceSize.height - 1) // investigate further if corners are equal if (topLeft == bottomLeft && topLeft == topRight && topLeft == bottomRight) { breakable { for (sx <- x until x + surfaceSize.width; sy <- y until y + surfaceSize.height) { if (matrix(sx)(sy) != topLeft) { x = if (x == sx) sx + 1 else sx break } } // found one! resultPositions ::= Position2D(x, y) x += 1 } } else if (topRight != bottomRight) { // can skip x a bit as there won't be a valid match in current row in this area x += surfaceSize.width } else { x += 1 } } } return resultPositions } I already tried to include some optimizations in it but I am sure that there are far better solutions. Is there a matlab function existing for it which I could port? I'm also wondering whether this problem has its own name as I didn't exactly know what to google for. Thanks for thinking about it! I'm excited to see your proposals or solutions :)

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  • Should the argument be passed by reference in this .net example?

    - by Hamish Grubijan
    I have used Java, C++, .Net. (in that order). When asked about by-value vs. by-ref on interviews, I have always done well on that question ... perhaps because nobody went in-depth on it. Now I know that I do not see the whole picture. I was looking at this section of code written by someone else: XmlDocument doc = new XmlDocument(); AppendX(doc); // Real name of the function is different AppendY(doc); // ditto When I saw this code, I thought: wait a minute, should not I use a ref in front of doc variable (and modify AppendX/Y accordingly? it works as written, but made me question whether I actually understand the ref keyword in C#. As I thought about this more, I recalled early Java days (college intro language). A friend of mine looked at some code I have written and he had a mental block - he kept asking me which things are passed in by reference and when by value. My ignorant response was something like: Dude, there is only one kind of arg passing in Java and I forgot which one it is :). Chill, do not over-think and just code. Java still does not have a ref does it? Yet, Java hackers seem to be productive. Anyhow, coding in C++ exposed me to this whole by reference business, and now I am confused. Should ref be used in the example above? I am guessing that when ref is applied to value types: primitives, enums, structures (is there anything else in this list?) it makes a big difference. And ... when applied to objects it does not because it is all by reference. If things were so simple, then why would not the compiler restrict the usage of ref keyword to a subset of types. When it comes to objects, does ref serve as a comment sort of? Well, I do remember that there can be problems with null and ref is also useful for initializing multiple elements within a method (since you cannot return multiple things with the same easy as you would do in Python). Thanks.

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  • How to give properties to c++ classes (interfaces)

    - by caas
    Hello, I have built several classes (A, B, C...) which perform operations on the same BaseClass. Example: struct BaseClass { int method1(); int method2(); int method3(); } struct A { int methodA(BaseClass& bc) { return bc.method1(); } } struct B { int methodB(BaseClass& bc) { return bc.method2()+bc.method1(); } } struct C { int methodC(BaseClass& bc) { return bc.method3()+bc.method2(); } } But as you can see, each class A, B, C... only uses a subset of the available methods of the BaseClass and I'd like to split the BaseClass into several chunks such that it is clear what it used and what is not. For example a solution could be to use multiple inheritance: // A uses only method1() struct InterfaceA { virtual int method1() = 0; } struct A { int methodA(InterfaceA&); } // B uses method1() and method2() struct InterfaceB { virtual int method1() = 0; virtual int method2() = 0; } struct B { int methodB(InterfaceB&); } // C uses method2() and method3() struct InterfaceC { virtual int method2() = 0; virtual int method3() = 0; } struct C { int methodC(InterfaceC&); } The problem is that each time I add a new type of operation, I need to change the implementation of BaseClass. For example: // D uses method1() and method3() struct InterfaceD { virtual int method1() = 0; virtual int method3() = 0; } struct D { int methodD(InterfaceD&); } struct BaseClass : public A, B, C // here I need to add class D { ... } Do you know a clean way I can do this? Thanks for your help edit: I forgot to mention that it can also be done with templates. But I don't like this solution either because the required interface does not appear explicitly in the code. You have to try to compile the code to verify that all required methods are implemented correctly. Plus, it would require to instantiate different versions of the classes (one for each BaseClass type template parameter) and this is not always possible nor desired.

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  • body onload cache not clearing

    - by Mad Cow
    I'm using an image swapping function generated by Dreamweaver to allow an image to change when moused over. The images are small. I have a problem because the images are getting stored in cache and without clearing it out I cant get the new images to show. It works on some browsers, but unfortunately not on all... I've read about putting "a random query" into the javascript to force the page to reload, but I dont know where to put it (the code was generated for me by dreamweaver). A subset of my code is : <script type="text/javascript"> function MM_preloadImages() { //v3.0 var d=document; if(d.images){ if(!d.MM_p) d.MM_p=new Array(); var i,j=d.MM_p.length,a=MM_preloadImages.arguments; for(i=0; i<a.length; i++) if (a[i].indexOf("#")!=0){ d.MM_p[j]=new Image; d.MM_p[j++].src=a[i];}} } function MM_swapImgRestore() { //v3.0 var i,x,a=document.MM_sr; for(i=0;a&&i<a.length&&(x=a[i])&&x.oSrc;i++) x.src=x.oSrc; } function MM_findObj(n, d) { //v4.01 var p,i,x; if(!d) d=document; if((p=n.indexOf("?"))>0&&parent.frames.length) { d=parent.frames[n.substring(p+1)].document; n=n.substring(0,p);} if(!(x=d[n])&&d.all) x=d.all[n]; for (i=0;!x&&i<d.forms.length;i++) x=d.forms[i][n]; for(i=0;!x&&d.layers&&i<d.layers.length;i++) x=MM_findObj(n,d.layers[i].document); if(!x && d.getElementById) x=d.getElementById(n); return x; } function MM_swapImage() { //v3.0 var i,j=0,x,a=MM_swapImage.arguments; document.MM_sr=new Array; for(i=0;i<(a.length-2);i+=3) if ((x=MM_findObj(a[i]))!=null){document.MM_sr[j++]=x; if(!x.oSrc) x.oSrc=x.src; x.src=a[i+2];} } </script> </head> <body onload="MM_preloadImages('../images/navigation/social-about-us-over.jpg','../images/navigation/social-about-us.jpg','../images/navigation/social-activities-over.jpg','../images/navigation/social-ourservices-over.jpg','../images/navigation/social-howwework-over.jpg','../images/navigation/social-fundraising-over.jpg','../images/navigation/social-howtohelp-over.jpg','../images/navigation/social-contactus-over.jpg')"> My website is http://www.clockhouse.org.uk/ I'm sure there is a better way i could have written this, but if anyone can help me fix this code I'd be very grateful Many thanks

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  • Sending Messages to SignalR Hubs from the Outside

    - by Ricardo Peres
    Introduction You are by now probably familiarized with SignalR, Microsoft’s API for real-time web functionality. This is, in my opinion, one of the greatest products Microsoft has released in recent time. Usually, people login to a site and enter some page which is connected to a SignalR hub. Then they can send and receive messages – not just text messages, mind you – to other users in the same hub. Also, the server can also take the initiative to send messages to all or a specified subset of users on its own, this is known as server push. The normal flow is pretty straightforward, Microsoft has done a great job with the API, it’s clean and quite simple to use. And for the latter – the server taking the initiative – it’s also quite simple, just involves a little more work. The Problem The API for sending messages can be achieved from inside a hub – an instance of the Hub class – which is something that we don’t have if we are the server and we want to send a message to some user or group of users: the Hub instance is only instantiated in response to a client message. The Solution It is possible to acquire a hub’s context from outside of an actual Hub instance, by calling GlobalHost.ConnectionManager.GetHubContext<T>(). This API allows us to: Broadcast messages to all connected clients (possibly excluding some); Send messages to a specific client; Send messages to a group of clients. So, we have groups and clients, each is identified by a string. Client strings are called connection ids and group names are free-form, given by us. The problem with client strings is, we do not know how these map to actual users. One way to achieve this mapping is by overriding the Hub’s OnConnected and OnDisconnected methods and managing the association there. Here’s an example: 1: public class MyHub : Hub 2: { 3: private static readonly IDictionary<String, ISet<String>> users = new ConcurrentDictionary<String, ISet<String>>(); 4:  5: public static IEnumerable<String> GetUserConnections(String username) 6: { 7: ISet<String> connections; 8:  9: users.TryGetValue(username, out connections); 10:  11: return (connections ?? Enumerable.Empty<String>()); 12: } 13:  14: private static void AddUser(String username, String connectionId) 15: { 16: ISet<String> connections; 17:  18: if (users.TryGetValue(username, out connections) == false) 19: { 20: connections = users[username] = new HashSet<String>(); 21: } 22:  23: connections.Add(connectionId); 24: } 25:  26: private static void RemoveUser(String username, String connectionId) 27: { 28: users[username].Remove(connectionId); 29: } 30:  31: public override Task OnConnected() 32: { 33: AddUser(this.Context.Request.User.Identity.Name, this.Context.ConnectionId); 34: return (base.OnConnected()); 35: } 36:  37: public override Task OnDisconnected() 38: { 39: RemoveUser(this.Context.Request.User.Identity.Name, this.Context.ConnectionId); 40: return (base.OnDisconnected()); 41: } 42: } As you can see, I am using a static field to store the mapping between a user and its possibly many connections – for example, multiple open browser tabs or even multiple browsers accessing the same page with the same login credentials. The user identity, as is normal in .NET, is obtained from the IPrincipal which in SignalR hubs case is stored in Context.Request.User. Of course, this property will only have a meaningful value if we enforce authentication. Another way to go is by creating a group for each user that connects: 1: public class MyHub : Hub 2: { 3: public override Task OnConnected() 4: { 5: this.Groups.Add(this.Context.ConnectionId, this.Context.Request.User.Identity.Name); 6: return (base.OnConnected()); 7: } 8:  9: public override Task OnDisconnected() 10: { 11: this.Groups.Remove(this.Context.ConnectionId, this.Context.Request.User.Identity.Name); 12: return (base.OnDisconnected()); 13: } 14: } In this case, we will have a one-to-one equivalence between users and groups. All connections belonging to the same user will fall in the same group. So, if we want to send messages to a user from outside an instance of the Hub class, we can do something like this, for the first option – user mappings stored in a static field: 1: public void SendUserMessage(String username, String message) 2: { 3: var context = GlobalHost.ConnectionManager.GetHubContext<MyHub>(); 4: 5: foreach (String connectionId in HelloHub.GetUserConnections(username)) 6: { 7: context.Clients.Client(connectionId).sendUserMessage(message); 8: } 9: } And for using groups, its even simpler: 1: public void SendUserMessage(String username, String message) 2: { 3: var context = GlobalHost.ConnectionManager.GetHubContext<MyHub>(); 4:  5: context.Clients.Group(username).sendUserMessage(message); 6: } Using groups has the advantage that the IHubContext interface returned from GetHubContext has direct support for groups, no need to send messages to individual connections. Of course, you can wrap both mapping options in a common API, perhaps exposed through IoC. One example of its interface might be: 1: public interface IUserToConnectionMappingService 2: { 3: //associate and dissociate connections to users 4:  5: void AddUserConnection(String username, String connectionId); 6:  7: void RemoveUserConnection(String username, String connectionId); 8: } SignalR has built-in dependency resolution, by means of the static GlobalHost.DependencyResolver property: 1: //for using groups (in the Global class) 2: GlobalHost.DependencyResolver.Register(typeof(IUserToConnectionMappingService), () => new GroupsMappingService()); 3:  4: //for using a static field (in the Global class) 5: GlobalHost.DependencyResolver.Register(typeof(IUserToConnectionMappingService), () => new StaticMappingService()); 6:  7: //retrieving the current service (in the Hub class) 8: var mapping = GlobalHost.DependencyResolver.Resolve<IUserToConnectionMappingService>(); Now all you have to do is implement GroupsMappingService and StaticMappingService with the code I shown here and change SendUserMessage method to rely in the dependency resolver for the actual implementation. Stay tuned for more SignalR posts!

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  • .NET 4.5 is an in-place replacement for .NET 4.0

    - by Rick Strahl
    With the betas for .NET 4.5 and Visual Studio 11 and Windows 8 shipping many people will be installing .NET 4.5 and hacking away on it. There are a number of great enhancements that are fairly transparent, but it's important to understand what .NET 4.5 actually is in terms of the CLR running on your machine. When .NET 4.5 is installed it effectively replaces .NET 4.0 on the machine. .NET 4.0 gets overwritten by a new version of .NET 4.5 which - according to Microsoft - is supposed to be 100% backwards compatible. While 100% backwards compatible sounds great, we all know that 100% is a hard number to hit, and even the aforementioned blog post at the Microsoft site acknowledges this. But there's so much more than backwards compatibility that makes this awkward at best and confusing at worst. What does ‘Replacement’ mean? When you install .NET 4.5 your .NET 4.0 assemblies in the \Windows\.NET Framework\V4.0.30319 are overwritten with a new set of assemblies. You end up with overwritten assemblies as well as a bunch of new ones (like the new System.Net.Http assemblies for example). The following screen shot demonstrates system.dll on my test machine (left) running .NET 4.5 on the right and my production laptop running stock .NET 4.0 (right):   Clearly they are different files with a difference in file sizes (interesting that the 4.5 version is actually smaller). That’s not all. If you actually query the runtime version when .NET 4.5 is installed with with Environment.Version you still get: 4.0.30319 If you open the properties of System.dll assembly in .NET 4.5 you'll also see: Notice that the file version is also left at 4.0.xxx. There are differences in build numbers: .NET 4.0 shows 261 and the current .NET 4.5 beta build is 17379. I suppose you can use assume a build number greater than 17000 is .NET 4.5, but that's pretty hokey to say the least. There’s no easy or obvious way to tell whether you are running on 4.0 or 4.5 – to the application they appear to be the same runtime version. And that is what Microsoft intends here. .NET 4.5 is intended as an in-place upgrade. Compile to 4.5 run on 4.0 – not quite! You can compile an application for .NET 4.5 and run it on the 4.0 runtime – that is until you hit a new feature that doesn’t exist on 4.0. At which point the app bombs at runtime. Say you write some code that is mostly .NET 4.0, but only has a few of the new features of .NET 4.5 like aync/await buried deep in the bowels of the application where it only fires occasionally. .NET will happily start your application and run everything 4.0 fine, until it hits that 4.5 code – and then crash unceremoniously at runtime. Oh joy! You can .NET 4.0 applications on .NET 4.5 of course and that should work without much fanfare. Different than .NET 3.0/3.5 Note that this in-place replacement is very different from the side by side installs of .NET 2.0 and 3.0/3.5 which all ran on the 2.0 version of the CLR. The two 3.x versions were basically library enhancements on top of the core .NET 2.0 runtime. Both versions ran under the .NET 2.0 runtime which wasn’t changed (other than for security patches and bug fixes) for the whole 3.x cycle. The 4.5 update instead completely replaces the .NET 4.0 runtime and leaves the actual version number set at v4.0.30319. When you build a new project with Visual Studio 2011, you can still target .NET 4.0 or you can target .NET 4.5. But you are in effect referencing the same set of assemblies for both regardless which version you use. What's different is the compiler used to compile and link your code so compiling with .NET 4.0 gives you just the subset of the functionality that is available in .NET 4.0, but when you use the 4.5 compiler you get the full functionality of what’s actually available in the assemblies and extra libraries. It doesn’t look like you will be able to use Visual Studio 2010 to develop .NET 4.5 applications. Good news – Bad news Microsoft is trying hard to experiment with every possible permutation of releasing new versions of the .NET framework apparently. No two updates have been the same. Clearly updating to a full new version of .NET (ie. .NET 2.0, 4.0 and at some point 5.0 runtimes) has its own set of challenges, but doing an in-place update of the runtime and then not even providing a good way to tell which version is installed is pretty whacky even by Microsoft’s standards. Especially given that .NET 4.5 includes a fairly significant update with all the aysnc functionality baked into the runtime. Most of the IO APIs have been updated to support task based async operation which significantly affects many existing APIs. To make things worse .NET 4.5 will be the initial version of .NET that ships with Windows 8 so it will be with us for a long time to come unless Microsoft finally decides to push .NET versions onto Windows machines as part of system upgrades (which currently doesn’t happen). This is the same story we had when Vista launched with .NET 3.0 which was a minor version that quickly was replaced by 3.5 which was more long lived and practical. People had enough problems dealing with the confusing versioning of the 3.x versions which ran on .NET 2.0. I can’t count the amount support calls and questions I’ve fielded because people couldn’t find a .NET 3.5 entry in the IIS version dialog. The same is likely to happen with .NET 4.5. It’s all well and good when we know that .NET 4.5 is an in-place replacement, but administrators and IT folks not intimately familiar with .NET are unlikely to understand this nuance and end up thoroughly confused which version is installed. It’s hard for me to see any upside to an in-place update and I haven’t really seen a good explanation of why this approach was decided on. Sure if the version stays the same existing assembly bindings don’t break so applications can stay running through an update. I suppose this is useful for some component vendors and strongly signed assemblies in corporate environments. But seriously, if you are going to throw .NET 4.5 into the mix, who won’t be recompiling all code and thoroughly test that code to work on .NET 4.5? A recompile requirement doesn’t seem that serious in light of a major version upgrade.  Resources http://blogs.msdn.com/b/dotnet/archive/2011/09/26/compatibility-of-net-framework-4-5.aspx http://www.devproconnections.com/article/net-framework/net-framework-45-versioning-faces-problems-141160© Rick Strahl, West Wind Technologies, 2005-2012Posted in .NET   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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