Search Results

Search found 21336 results on 854 pages for 'db api'.

Page 259/854 | < Previous Page | 255 256 257 258 259 260 261 262 263 264 265 266  | Next Page >

  • Error accessing a Web Service with SSL

    - by Elie
    I have a program that is supposed to send a file to a web service, which requires an SSL connection. I run the program as follows: SET JAVA_HOME=C:\Program Files\Java\jre1.6.0_07 SET com.ibm.SSL.ConfigURL=ssl.client.props "%JAVA_HOME%\bin\java" -cp ".;Test.jar" ca.mypackage.Main This was works fine, but when I change the first line to SET JAVA_HOME=C:\Program Files\IBM\SDP\runtimes\base_v7\java\jre I get the following error: com.sun.xml.internal.ws.client.ClientTransportException: HTTP transport error: java.net.SocketException: java.lang.ClassNotFoundException: Cannot find the specified class com.ibm.websphere.ssl.protocol.SSLSocketFactory at com.sun.xml.internal.ws.transport.http.client.HttpClientTransport.getOutput(HttpClientTransport.java:119) at com.sun.xml.internal.ws.transport.http.client.HttpTransportPipe.process(HttpTransportPipe.java:140) at com.sun.xml.internal.ws.transport.http.client.HttpTransportPipe.processRequest(HttpTransportPipe.java:86) at com.sun.xml.internal.ws.api.pipe.Fiber.__doRun(Fiber.java:593) at com.sun.xml.internal.ws.api.pipe.Fiber._doRun(Fiber.java:552) at com.sun.xml.internal.ws.api.pipe.Fiber.doRun(Fiber.java:537) at com.sun.xml.internal.ws.api.pipe.Fiber.runSync(Fiber.java:434) at com.sun.xml.internal.ws.client.Stub.process(Stub.java:247) at com.sun.xml.internal.ws.client.sei.SEIStub.doProcess(SEIStub.java:132) at com.sun.xml.internal.ws.client.sei.SyncMethodHandler.invoke(SyncMethodHandler.java:242) at com.sun.xml.internal.ws.client.sei.SyncMethodHandler.invoke(SyncMethodHandler.java:222) at com.sun.xml.internal.ws.client.sei.SEIStub.invoke(SEIStub.java:115) at $Proxy26.fileSubmit(Unknown Source) at com.testing.TestingSoapProxy.fileSubmit(TestingSoapProxy.java:81) at ca.mypackage.Main.main(Main.java:63) Caused by: java.net.SocketException: java.lang.ClassNotFoundException: Cannot find the specified class com.ibm.websphere.ssl.protocol.SSLSocketFactory at javax.net.ssl.DefaultSSLSocketFactory.a(SSLSocketFactory.java:7) at javax.net.ssl.DefaultSSLSocketFactory.createSocket(SSLSocketFactory.java:1) at com.ibm.net.ssl.www2.protocol.https.c.afterConnect(c.java:110) at com.ibm.net.ssl.www2.protocol.https.d.connect(d.java:14) at sun.net.www.protocol.http.HttpURLConnection.getOutputStream(HttpURLConnection.java:902) at com.ibm.net.ssl.www2.protocol.https.b.getOutputStream(b.java:86) at com.sun.xml.internal.ws.transport.http.client.HttpClientTransport.getOutput(HttpClientTransport.java:107) ... 14 more Caused by: java.lang.ClassNotFoundException: Cannot find the specified class com.ibm.websphere.ssl.protocol.SSLSocketFactory at javax.net.ssl.SSLJsseUtil.b(SSLJsseUtil.java:20) at javax.net.ssl.SSLSocketFactory.getDefault(SSLSocketFactory.java:36) at javax.net.ssl.HttpsURLConnection.getDefaultSSLSocketFactory(HttpsURLConnection.java:16) at javax.net.ssl.HttpsURLConnection.<init>(HttpsURLConnection.java:36) at com.ibm.net.ssl.www2.protocol.https.b.<init>(b.java:1) at com.ibm.net.ssl.www2.protocol.https.Handler.openConnection(Handler.java:11) at java.net.URL.openConnection(URL.java:995) at com.sun.xml.internal.ws.api.EndpointAddress.openConnection(EndpointAddress.java:206) at com.sun.xml.internal.ws.transport.http.client.HttpClientTransport.createHttpConnection(HttpClientTransport.java:277) at com.sun.xml.internal.ws.transport.http.client.HttpClientTransport.getOutput(HttpClientTransport.java:103) ... 14 more So it seems that this problem would be related to the JRE I'm using, but what doesn't seem to make sense is that the non-IBM JRE works fine, but the IBM JRE does not. Any ideas, or suggestions?

    Read the article

  • Redirect output logs of javax.xml.ws and com.sun.xml.ws

    - by chrisnfoneur
    I am working on a SOAP based web service, with Sun's Metro. I am facing an annoying bug, each time I send a malformed SOAP object to my web service, sun's api spam the System.out with logs like this: javax.xml.ws.WebServiceException: com.sun.istack.XMLStreamException2: org.xml.sax.SAXParseException: cvc-complex-type.4: Attribute 'type' must appear on element 'object'. at com.sun.xml.ws.util.pipe.AbstractSchemaValidationTube.doProcess(AbstractSchemaValidationTube.java:206) at com.sun.xml.ws.util.pipe.AbstractSchemaValidationTube.processRequest(AbstractSchemaValidationTube.java:175) at com.sun.xml.ws.api.pipe.Fiber.__doRun(Fiber.java:595) at com.sun.xml.ws.api.pipe.Fiber._doRun(Fiber.java:554) at com.sun.xml.ws.api.pipe.Fiber.doRun(Fiber.java:539) at com.sun.xml.ws.api.pipe.Fiber.runSync(Fiber.java:436) at com.sun.xml.ws.server.WSEndpointImpl$2.process(WSEndpointImpl.java:243) at com.sun.xml.ws.transport.http.HttpAdapter$HttpToolkit.handle(HttpAdapter.java:444) at com.sun.xml.ws.transport.http.HttpAdapter.handle(HttpAdapter.java:244) at com.sun.xml.ws.transport.http.server.WSHttpHandler.handleExchange(WSHttpHandler.java:106) at com.sun.xml.ws.transport.http.server.WSHttpHandler.handle(WSHttpHandler.java:91) at com.sun.net.httpserver.Filter$Chain.doFilter(Filter.java:65) at sun.net.httpserver.AuthFilter.doFilter(AuthFilter.java:54) at com.sun.net.httpserver.Filter$Chain.doFilter(Filter.java:68) at sun.net.httpserver.ServerImpl$Exchange$LinkHandler.handle(ServerImpl.java:555) at com.sun.net.httpserver.Filter$Chain.doFilter(Filter.java:65) at sun.net.httpserver.ServerImpl$Exchange.run(ServerImpl.java:527) at sun.net.httpserver.ServerImpl$DefaultExecutor.execute(ServerImpl.java:119) at sun.net.httpserver.ServerImpl$Dispatcher.handle(ServerImpl.java:349) at sun.net.httpserver.ServerImpl$Dispatcher.run(ServerImpl.java:321) at java.lang.Thread.run(Thread.java:619) Caused by: com.sun.istack.XMLStreamException2: org.xml.sax.SAXParseException: cvc-complex-type.4: Attribute 'type' must appear on element 'object'. at com.sun.xml.ws.util.xml.StAXSource$1.parse(StAXSource.java:185) at com.sun.xml.ws.util.xml.StAXSource$1.parse(StAXSource.java:170) at org.apache.xerces.jaxp.validation.ValidatorHandlerImpl.validate(Unknown Source) at org.apache.xerces.jaxp.validation.ValidatorImpl.validate(Unknown Source) at javax.xml.validation.Validator.validate(Validator.java:127) at com.sun.xml.ws.util.pipe.AbstractSchemaValidationTube.doProcess(AbstractSchemaValidationTube.java:204) ... 20 more I would like to switch off this log or redirect it to my error/warn/debug.log files used by log4j. I tried to add a rule in my log4j.xml file : <category name="javax.xml.ws"> <priority value="error" /> </category> It didn't worked. So I tried the following trick: java.util.logging.Logger.getLogger("javax.xml.ws.WebServiceException").setLevel( java.util.logging.Level.OFF); it didn't worked neither. Any ideas ? It is not a big issue but it makes my catalina.log getting bigger & bigger and it's not the appropriate place for this kind of log. Chris

    Read the article

  • Help with OpenSSL request using Python

    - by Ldn
    Hi i'm creating a program that has to make a request and then obtain some info. For doing that the website had done some API that i will use. There is an how-to about these API but every example is made using PHP. But my app is done using Python so i need to convert the code. here is the how-to: The request string is sealed with OpenSSL. The steps for sealing are as follows: • Random 128-bit key is created. • Random key is used to RSA-RC4 symettrically encrypt the request string. • Random key is encrypted with the public key using OpenSSL RSA asymmetrical encryption. • The encrypted request and encrypted key are each base64 encoded and placed in the appropriate fields. In PHP a full request to our API can be accomplished like so: <?php // initial request. $request = array('object' => 'Link', 'action' => 'get', 'args' => array( 'app_id' => 303612602 ) ); // encode the request in JSON $request = json_encode($request); // when you receive your profile, you will be given a public key to seal your request in. $key_pem = "-----BEGIN PUBLIC KEY----- MFwwDQYJKoZIhvcNAQEBBQADSwAwSAJBALdu5C6d2sA1Lu71NNGBEbLD6DjwhFQO VLdFAJf2rOH63rG/L78lrQjwMLZOeHEHqjaiUwCr8NVTcVrebu6ylIECAwEAAQ== -----END PUBLIC KEY-----"; // load the public key $pkey = openssl_pkey_get_public($key_pem); // seal! $newrequest and $enc_keys are passed by reference. openssl_seal($request, $enc_request, $enc_keys, array($pkey)); // then wrap the request $wrapper = array( 'profile' => 'ProfileName', 'format' => 'RSA_RC4_Sealed', 'enc_key' => base64_encode($enc_keys[0]), 'request' => base64_encode($enc_request) ); // json encode the wrapper. urlencode it as well. $wrapper = urlencode(json_encode($wrapper)); // we can send the request wrapper via the cURL extension $ch = curl_init(); curl_setopt($ch, CURLOPT_URL, 'http://api.site.com/'); curl_setopt($ch, CURLOPT_POST, 1); curl_setopt($ch, CURLOPT_POSTFIELDS, "request=$wrapper"); curl_setopt($ch, CURLOPT_RETURNTRANSFER, true); $data = curl_exec($ch); curl_close($ch); ?> Of all of that, i was able to convert "$request" and i'v also made the JSON encode. This is my code: import urllib import urllib2 import json url = 'http://api.site.com/' array = {'app_id' : "303612602"} values = { "object" : "Link", "action": "get", "args" : array } data = urllib.urlencode(values) json_data = json.dumps(data) What stop me is the sealing with OpenSSL and the publi key (that obviously i have) Using PHP OpenSSL it's so easy, but in Python i don't really know how to use it Please, help me!

    Read the article

  • jQuery .ajax call to bit.ly returns results in IE but not FF or Chrome

    - by Ian Quigley
    I am trying to call to the bit.ly URL shortening service using jQuery with an .ajax call. <html><head> <script type="text/javascript" src="http://www.twipler.com/settings/scripts/jquery.1.4.min.js"></script> <script type="text/javascript"> jQuery.fn.shorten = function(url) { var resultUrl = url; $.ajax( { url: "http://api.bit.ly/shorten?version=2.0.1&login=twipler&apiKey=R_4e618e42fadbb802cf95c6c2dbab3763&longUrl=" + url, async: false, dataType: 'json', data: "", type: "GET", success: function (json) { resultUrl = json.results[url].shortUrl; } }); return resultUrl; } ; </script></head><body> <a href="#" onclick="alert($().shorten('http://amiconnectedtotheinternet.com'));"> Shorten</a> </body> </html> This works in IE8 but does not work in FireFox (3.5.9) nor in Chrome. In both cases 'json' is null. Headers in IE8 GET http://api.bit.ly/shorten?ver..[SNIP]..dtotheinternet.com HTTP/1.1 Accept: application/json, text/javascript, */* Accept-Language: en-US Accept-Encoding: gzip, deflate User-Agent: Mozilla/4.0 (compatible; MSIE 8.0; Windows NT 6.1; WOW64; Trident/4.0; SLCC2; .NET CLR 2.0.50727; Media Center PC 6.0; InfoPath.2; .NET CLR 1.1.4322; .NET CLR 3.5.30729; .NET CLR 3.0.30729) Host: api.bit.ly Connection: Keep-Alive Headers in Chrome GET http://api.bit.ly/shorten?versio..[SNIP]..nectedtotheinternet.com HTTP/1.1 Host: api.bit.ly Connection: keep-alive User-Agent: Mozilla/5.0 (Windows; U; Windows NT 6.1; en-US) AppleWebKit/532.5 (KHTML, like Gecko) Chrome/4.1.249.1045 Safari/532.5 Origin: file:// Accept: application/json, text/javascript, */* Accept-Encoding: gzip,deflate,sdch Accept-Language: en-US,en;q=0.8 Accept-Charset: ISO-8859-1,utf-8;q=0.7,*;q=0.3 So the only obvious difference is that Chrome is sending "Origin: file://" and I've no idea how to stop it doing that.

    Read the article

  • C#: BackgroundWorker cloning resources?

    - by Dav
    The problem I've been struggling with this partiular problem for two days now and just run out of ideas. A little... background: we have a WinForms app that needs to access a database, construct a list of related in-memory objects from that data, and then display on a DataGridView. Important point is that we first populate an app-wide cache (List), and then create a mirror of the cache local to the form on which the DGV lives (using List constructor param). Because fetching the data takes a good few seconds (DB sits on a LAN server) to load, we decided to use a BackgroundWorker, and only refresh the DGV once the data is loaded. However, it seems that doing the loading via a BGW results in some memory leak... or an error on my part. When loaded using a blocking method call, the app consumes about 30MB of RAM; with a BGW this jumps to 80MB! While it may not seem as much anyway, our clients are not too happy about it. Relevant code Form private void MyForm_Load(object sender, EventArgs e) { MyRepository.Instance.FinishedEvent += RefreshCache; } private void RefreshCache(object sender, EventArgs e) { dgvProducts.DataSource = new List<MyDataObj>(MyRepository.Products); } Repository private static List<MyDataObj> Products { get; set; } public event EventHandler ProductsLoaded; public void GetProductsSync() { List<MyDataObj> p; using (MyL2SDb db = new MyL2SDb(MyConfig.ConnectionString)) { p = db.PRODUCTS .Select(p => new MyDataObj {Id = p.ID, Description = p.DESCR}) .ToList(); } Products = p; // tell the form to refresh UI if (ProductsLoaded != null) ProductsLoaded(this, null); } public void GetProductsAsync() { using (BackgroundWorker myWorker = new BackgroundWorker()) { myWorker.DoWork += delegate { List<MyDataObj> p; using (MyL2SDb db = new MyL2SDb(MyConfig.ConnectionString)) { p = db.PRODUCTS .Select(p => new MyDataObj {Id = p.ID, Description = p.DESCR}) .ToList(); } Products = p; }; // tell the form to refresh UI when finished myWorker.RunWorkerCompleted += GetProductsCompleted; myWorker.RunWorkerAsync(); } } private void GetProductsCompleted(object sender, RunWorkerCompletedEventArgs e) { if (ProductsLoaded != null) ProductsLoaded(this, null); } End! GetProductsSync or GetProductsAsync are called on the main thread, not shown above. Could it be that the GarbageCollector just gets lost with two threads? Or is it the task manager that shows incorrect values? Will be greateful for any responses, suggestions, criticism.

    Read the article

  • Informed TDD &ndash; Kata &ldquo;To Roman Numerals&rdquo;

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/05/28/informed-tdd-ndash-kata-ldquoto-roman-numeralsrdquo.aspxIn a comment on my article on what I call Informed TDD (ITDD) reader gustav asked how this approach would apply to the kata “To Roman Numerals”. And whether ITDD wasn´t a violation of TDD´s principle of leaving out “advanced topics like mocks”. I like to respond with this article to his questions. There´s more to say than fits into a commentary. Mocks and TDD I don´t see in how far TDD is avoiding or opposed to mocks. TDD and mocks are orthogonal. TDD is about pocess, mocks are about structure and costs. Maybe by moving forward in tiny red+green+refactor steps less need arises for mocks. But then… if the functionality you need to implement requires “expensive” resource access you can´t avoid using mocks. Because you don´t want to constantly run all your tests against the real resource. True, in ITDD mocks seem to be in almost inflationary use. That´s not what you usually see in TDD demonstrations. However, there´s a reason for that as I tried to explain. I don´t use mocks as proxies for “expensive” resource. Rather they are stand-ins for functionality not yet implemented. They allow me to get a test green on a high level of abstraction. That way I can move forward in a top-down fashion. But if you think of mocks as “advanced” or if you don´t want to use a tool like JustMock, then you don´t need to use mocks. You just need to stand the sight of red tests for a little longer ;-) Let me show you what I mean by that by doing a kata. ITDD for “To Roman Numerals” gustav asked for the kata “To Roman Numerals”. I won´t explain the requirements again. You can find descriptions and TDD demonstrations all over the internet, like this one from Corey Haines. Now here is, how I would do this kata differently. 1. Analyse A demonstration of TDD should never skip the analysis phase. It should be made explicit. The requirements should be formalized and acceptance test cases should be compiled. “Formalization” in this case to me means describing the API of the required functionality. “[D]esign a program to work with Roman numerals” like written in this “requirement document” is not enough to start software development. Coding should only begin, if the interface between the “system under development” and its context is clear. If this interface is not readily recognizable from the requirements, it has to be developed first. Exploration of interface alternatives might be in order. It might be necessary to show several interface mock-ups to the customer – even if that´s you fellow developer. Designing the interface is a task of it´s own. It should not be mixed with implementing the required functionality behind the interface. Unfortunately, though, this happens quite often in TDD demonstrations. TDD is used to explore the API and implement it at the same time. To me that´s a violation of the Single Responsibility Principle (SRP) which not only should hold for software functional units but also for tasks or activities. In the case of this kata the API fortunately is obvious. Just one function is needed: string ToRoman(int arabic). And it lives in a class ArabicRomanConversions. Now what about acceptance test cases? There are hardly any stated in the kata descriptions. Roman numerals are explained, but no specific test cases from the point of view of a customer. So I just “invent” some acceptance test cases by picking roman numerals from a wikipedia article. They are supposed to be just “typical examples” without special meaning. Given the acceptance test cases I then try to develop an understanding of the problem domain. I´ll spare you that. The domain is trivial and is explain in almost all kata descriptions. How roman numerals are built is not difficult to understand. What´s more difficult, though, might be to find an efficient solution to convert into them automatically. 2. Solve The usual TDD demonstration skips a solution finding phase. Like the interface exploration it´s mixed in with the implementation. But I don´t think this is how it should be done. I even think this is not how it really works for the people demonstrating TDD. They´re simplifying their true software development process because they want to show a streamlined TDD process. I doubt this is helping anybody. Before you code you better have a plan what to code. This does not mean you have to do “Big Design Up-Front”. It just means: Have a clear picture of the logical solution in your head before you start to build a physical solution (code). Evidently such a solution can only be as good as your understanding of the problem. If that´s limited your solution will be limited, too. Fortunately, in the case of this kata your understanding does not need to be limited. Thus the logical solution does not need to be limited or preliminary or tentative. That does not mean you need to know every line of code in advance. It just means you know the rough structure of your implementation beforehand. Because it should mirror the process described by the logical or conceptual solution. Here´s my solution approach: The arabic “encoding” of numbers represents them as an ordered set of powers of 10. Each digit is a factor to multiply a power of ten with. The “encoding” 123 is the short form for a set like this: {1*10^2, 2*10^1, 3*10^0}. And the number is the sum of the set members. The roman “encoding” is different. There is no base (like 10 for arabic numbers), there are just digits of different value, and they have to be written in descending order. The “encoding” XVI is short for [10, 5, 1]. And the number is still the sum of the members of this list. The roman “encoding” thus is simpler than the arabic. Each “digit” can be taken at face value. No multiplication with a base required. But what about IV which looks like a contradiction to the above rule? It is not – if you accept roman “digits” not to be limited to be single characters only. Usually I, V, X, L, C, D, M are viewed as “digits”, and IV, IX etc. are viewed as nuisances preventing a simple solution. All looks different, though, once IV, IX etc. are taken as “digits”. Then MCMLIV is just a sum: M+CM+L+IV which is 1000+900+50+4. Whereas before it would have been understood as M-C+M+L-I+V – which is more difficult because here some “digits” get subtracted. Here´s the list of roman “digits” with their values: {1, I}, {4, IV}, {5, V}, {9, IX}, {10, X}, {40, XL}, {50, L}, {90, XC}, {100, C}, {400, CD}, {500, D}, {900, CM}, {1000, M} Since I take IV, IX etc. as “digits” translating an arabic number becomes trivial. I just need to find the values of the roman “digits” making up the number, e.g. 1954 is made up of 1000, 900, 50, and 4. I call those “digits” factors. If I move from the highest factor (M=1000) to the lowest (I=1) then translation is a two phase process: Find all the factors Translate the factors found Compile the roman representation Translation is just a look-up. Finding, though, needs some calculation: Find the highest remaining factor fitting in the value Remember and subtract it from the value Repeat with remaining value and remaining factors Please note: This is just an algorithm. It´s not code, even though it might be close. Being so close to code in my solution approach is due to the triviality of the problem. In more realistic examples the conceptual solution would be on a higher level of abstraction. With this solution in hand I finally can do what TDD advocates: find and prioritize test cases. As I can see from the small process description above, there are two aspects to test: Test the translation Test the compilation Test finding the factors Testing the translation primarily means to check if the map of factors and digits is comprehensive. That´s simple, even though it might be tedious. Testing the compilation is trivial. Testing factor finding, though, is a tad more complicated. I can think of several steps: First check, if an arabic number equal to a factor is processed correctly (e.g. 1000=M). Then check if an arabic number consisting of two consecutive factors (e.g. 1900=[M,CM]) is processed correctly. Then check, if a number consisting of the same factor twice is processed correctly (e.g. 2000=[M,M]). Finally check, if an arabic number consisting of non-consecutive factors (e.g. 1400=[M,CD]) is processed correctly. I feel I can start an implementation now. If something becomes more complicated than expected I can slow down and repeat this process. 3. Implement First I write a test for the acceptance test cases. It´s red because there´s no implementation even of the API. That´s in conformance with “TDD lore”, I´d say: Next I implement the API: The acceptance test now is formally correct, but still red of course. This will not change even now that I zoom in. Because my goal is not to most quickly satisfy these tests, but to implement my solution in a stepwise manner. That I do by “faking” it: I just “assume” three functions to represent the transformation process of my solution: My hypothesis is that those three functions in conjunction produce correct results on the API-level. I just have to implement them correctly. That´s what I´m trying now – one by one. I start with a simple “detail function”: Translate(). And I start with all the test cases in the obvious equivalence partition: As you can see I dare to test a private method. Yes. That´s a white box test. But as you´ll see it won´t make my tests brittle. It serves a purpose right here and now: it lets me focus on getting one aspect of my solution right. Here´s the implementation to satisfy the test: It´s as simple as possible. Right how TDD wants me to do it: KISS. Now for the second equivalence partition: translating multiple factors. (It´a pattern: if you need to do something repeatedly separate the tests for doing it once and doing it multiple times.) In this partition I just need a single test case, I guess. Stepping up from a single translation to multiple translations is no rocket science: Usually I would have implemented the final code right away. Splitting it in two steps is just for “educational purposes” here. How small your implementation steps are is a matter of your programming competency. Some “see” the final code right away before their mental eye – others need to work their way towards it. Having two tests I find more important. Now for the next low hanging fruit: compilation. It´s even simpler than translation. A single test is enough, I guess. And normally I would not even have bothered to write that one, because the implementation is so simple. I don´t need to test .NET framework functionality. But again: if it serves the educational purpose… Finally the most complicated part of the solution: finding the factors. There are several equivalence partitions. But still I decide to write just a single test, since the structure of the test data is the same for all partitions: Again, I´m faking the implementation first: I focus on just the first test case. No looping yet. Faking lets me stay on a high level of abstraction. I can write down the implementation of the solution without bothering myself with details of how to actually accomplish the feat. That´s left for a drill down with a test of the fake function: There are two main equivalence partitions, I guess: either the first factor is appropriate or some next. The implementation seems easy. Both test cases are green. (Of course this only works on the premise that there´s always a matching factor. Which is the case since the smallest factor is 1.) And the first of the equivalence partitions on the higher level also is satisfied: Great, I can move on. Now for more than a single factor: Interestingly not just one test becomes green now, but all of them. Great! You might say, then I must have done not the simplest thing possible. And I would reply: I don´t care. I did the most obvious thing. But I also find this loop very simple. Even simpler than a recursion of which I had thought briefly during the problem solving phase. And by the way: Also the acceptance tests went green: Mission accomplished. At least functionality wise. Now I´ve to tidy up things a bit. TDD calls for refactoring. Not uch refactoring is needed, because I wrote the code in top-down fashion. I faked it until I made it. I endured red tests on higher levels while lower levels weren´t perfected yet. But this way I saved myself from refactoring tediousness. At the end, though, some refactoring is required. But maybe in a different way than you would expect. That´s why I rather call it “cleanup”. First I remove duplication. There are two places where factors are defined: in Translate() and in Find_factors(). So I factor the map out into a class constant. Which leads to a small conversion in Find_factors(): And now for the big cleanup: I remove all tests of private methods. They are scaffolding tests to me. They only have temporary value. They are brittle. Only acceptance tests need to remain. However, I carry over the single “digit” tests from Translate() to the acceptance test. I find them valuable to keep, since the other acceptance tests only exercise a subset of all roman “digits”. This then is my final test class: And this is the final production code: Test coverage as reported by NCrunch is 100%: Reflexion Is this the smallest possible code base for this kata? Sure not. You´ll find more concise solutions on the internet. But LOC are of relatively little concern – as long as I can understand the code quickly. So called “elegant” code, however, often is not easy to understand. The same goes for KISS code – especially if left unrefactored, as it is often the case. That´s why I progressed from requirements to final code the way I did. I first understood and solved the problem on a conceptual level. Then I implemented it top down according to my design. I also could have implemented it bottom-up, since I knew some bottom of the solution. That´s the leaves of the functional decomposition tree. Where things became fuzzy, since the design did not cover any more details as with Find_factors(), I repeated the process in the small, so to speak: fake some top level, endure red high level tests, while first solving a simpler problem. Using scaffolding tests (to be thrown away at the end) brought two advantages: Encapsulation of the implementation details was not compromised. Naturally private methods could stay private. I did not need to make them internal or public just to be able to test them. I was able to write focused tests for small aspects of the solution. No need to test everything through the solution root, the API. The bottom line thus for me is: Informed TDD produces cleaner code in a systematic way. It conforms to core principles of programming: Single Responsibility Principle and/or Separation of Concerns. Distinct roles in development – being a researcher, being an engineer, being a craftsman – are represented as different phases. First find what, what there is. Then devise a solution. Then code the solution, manifest the solution in code. Writing tests first is a good practice. But it should not be taken dogmatic. And above all it should not be overloaded with purposes. And finally: moving from top to bottom through a design produces refactored code right away. Clean code thus almost is inevitable – and not left to a refactoring step at the end which is skipped often for different reasons.   PS: Yes, I have done this kata several times. But that has only an impact on the time needed for phases 1 and 2. I won´t skip them because of that. And there are no shortcuts during implementation because of that.

    Read the article

  • SaaS Architecture Question from Newbie

    - by user226767
    I have developed a number of departmental client-server applications, and am now ready to begin working on moving one of these applications to a SaaS model. I have done some basic web development, but I'm a newbie when it comes to SaaS architectures. One of the first questions that comes to mind as I try to design the architecture is the question of single vs. multi tenancy. The pros and cons of each vary significantly depending on the type of application and scale required, so I'd like to describe my application and scale needs below, and hope others can comment on how I should get started with the architecture. The client-server application currently consists of a Firebird database and a Windows application. The database contains about 20 tables containing a few thousand records in 4 primary tables, and a few hundred records in various lookup and related tables. Although the number of records is small, the size can get large, as the database can contain large BLOBS. Each customer sets up their own database and has a handful of users within the organization connected to it. When I update the db schema, a new windows application is released, and it checks the db schema and then applies the updates as needed. For the SaaS application, I am designing for 100's (not 1000's or millions) of new customers per year. My first thought was to go with a multi tenancy model to make updates easy (shut down apply the updates to one database, and then start up). On the other hand, a single tenancy model would provide a means to roll updates out to a group of customers at a time, and spread the risk of data corruption - i.e. if something goes wrong with a database, it will impact one customer instead of all customers. With this idea, I was thinking of having a single web front-end which would connect to a single customer database upon login. Thus, when a new customer creates an account, a new database would be created (each customer would have their own db with multiple users as needed for the customer). In this model, a db update would require either a process to go through each db to apply schema changes, or a trigger upon logging in to initiate a schema update similar to the client-server model currently in use. Can anyone point me to information for similar applications which have been ported from client-server to SaaS? Or provide any pointers to consider? Basically I'm looking for architecture examples of taking a departmental application and making it available as a self service website for multiple customers. Thanks for any suggestions, resources, etc.

    Read the article

  • What are the benefits of using ORM over XML Serialization/Deserialization?

    - by Tequila Jinx
    I've been reading about NHibernate and Microsoft's Entity Framework to perform Object Relational Mapping against my data access layer. I'm interested in the benefits of having an established framework to perform ORM, but I'm curious as to the performance costs of using it against standard XML Serialization and Deserialization. Right now, I develop stored procedures in Oracle and SQL Server that use XML Types for either input or output parameters and return or shred XML depending on need. I use a custom database command object that uses generics to deserialize the XML results into a specified serializable class. By using a combination of generics, xml (de)serialization and Microsoft's DAAB, I've got a process that's fairly simple to develop against regardless of the data source. Moreover, since I exclusively use Stored Procedures to perform database operations, I'm mostly protected from changes in the data structure. Here's an over-simplified example of what I've been doing. static void main() { testXmlClass test = new test(1); test.Name = "Foo"; test.Save(); } // Example Serializable Class ------------------------------------------------ [XmlRootAttribute("test")] class testXmlClass() { [XmlElement(Name="id")] public int ID {set; get;} [XmlElement(Name="name")] public string Name {set; get;} //create an instance of the class loaded with data. public testXmlClass(int id) { GenericDBProvider db = new GenericDBProvider(); this = db.ExecuteSerializable("myGetByIDProcedure"); } //save the class to the database... public Save() { GenericDBProvider db = new GenericDBProvider(); db.AddInParameter("myInputParameter", DbType.XML, this); db.ExecuteSerializableNonQuery("mySaveProcedure"); } } // Database Handler ---------------------------------------------------------- class GenericDBProvider { public T ExecuteSerializable<T>(string commandText) where T : class { XmlSerializer xml = new XmlSerializer(typeof(T)); // connection and command code is assumed for the purposes of this example. // the final results basically just come down to... return xml.Deserialize(commandResults) as T; } public void ExecuteSerializableNonQuery(string commandText) { // once again, connection and command code is assumed... // basically, just execute the command along with the specified // parameters which have been serialized. } public void AddInParameter(string name, DbType type, object value) { StringWriter w = new StringWriter(); XmlSerializer x = new XmlSerializer(value.GetType()); //handle serialization for serializable classes. if (type == DbType.Xml && (value.GetType() != typeof(System.String))) { x.Serialize(w, value); w.Close(); // store serialized object in a DbParameterCollection accessible // to my other methods. } else { //handle all other parameter types } } } I'm starting a new project which will rely heavily on database operations. I'm very curious to know whether my current practices will be sustainable in a high-traffic situation and whether or not I should consider switching to NHibernate or Microsoft's Entity Framework to perform what essentially seems to boil down to the same thing I'm currently doing. I appreciate any advice you may have.

    Read the article

  • php , SimpleXML, while loop

    - by Michael
    I'm trying to get some information from ebay api and store it in database . I used simple xml to extract the information but I have a small issue as the information is not displayed for some items . if I make a print to the simple_xml I can see very well that the information is provided by ebay api . I have $items = "220617293997,250645537939,230485306218,110537213815,180519294810"; $number_of_items = count(explode(",", $items)); $xml = $baseClass->getContent("http://open.api.ebay.com/shopping?callname=GetMultipleItems&responseencoding=XML&appid=Morcovar-c74b-47c0-954f-463afb69a4b3&siteid=0&version=525&IncludeSelector=ItemSpecifics&ItemID=$items"); writeDoc($xml, "api.xml"); //echo $xml; $getvalues = simplexml_load_file('api.xml'); // print_r($getvalue); $number = "0"; while($number < 6) { $item_number = $getvalues->Item[$number]->ItemID; $location = $getvalues->Item[$number]->Location; $title = $getvalues->Item[$number]->Title; $price = $getvalues->Item[$number]->ConvertedCurrentPrice; $manufacturer = $getvalues->Item[$number]->ItemSpecifics->NameValueList[3]->Value; $model = $getvalues->Item[$number]->ItemSpecifics->NameValueList[4]->Value; $mileage = $getvalues->Item[$number]->ItemSpecifics->NameValueList[5]->Value; echo "item number = $item_number <br>localtion = $location<br>". "title = $title<br>price = $price<br>manufacturer = $manufacturer". "<br>model = $model<br>mileage = $mileage<br>"; $number++; } the above code returns item number = localtion = title = price = manufacturer = model = mileage = item number = 230485306218 localtion = Coventry, Warwickshire title = 2001 LAND ROVER RANGE ROVER VOGUE AUTO GREEN price = 3635.07 manufacturer = Land Rover model = Range Rover mileage = 76000 item number = 220617293997 localtion = Crawley, West Sussex title = 2004 CITROEN C5 HDI LX RED price = 3115.77 manufacturer = Citroen model = C5 mileage = 76000 item number = 180519294810 localtion = London, London title = 2000 VOLKSWAGEN POLO 1.4 SILVER 16V NEED GEAR BOX price = 905.06 manufacturer = Right-hand drive model = mileage = Standard Car item number = localtion = title = price = manufacturer = model = mileage = As you can see the information is not retrieved for a few items ... If I replace the $number manually like " $item_number = $getvalues-Item[4]-ItemID;" works well for any number .

    Read the article

  • The question about the basics of LINQ to SQL

    - by Alex
    I just started learning LINQ to SQL, and so far I'm impressed with the easy of use and good performance. I used to think that when doing LINQ queries like from Customer in DB.Customers where Customer.Age > 30 select Customer LINQ gets all customers from the database ("SELECT * FROM Customers"), moves them to the Customers array and then makes a search in that Array using .NET methods. This is very inefficient, what if there are hundreds of thousands of customers in the database? Making such big SELECT queries would kill the web application. Now after experiencing how actually fast LINQ to SQL is, I start to suspect that when doing that query I just wrote, LINQ somehow converts it to a SQL Query string SELECT * FROM Customers WHERE Age > 30 And only when necessary it will run the query. So my question is: am I right? And when is the query actually run? The reason why I'm asking is not only because I want to understand how it works in order to build good optimized applications, but because I came across the following problem. I have 2 tables, one of them is Books, the other has information on how many books were sold on certain days. My goal is to select books that had at least 50 sales/day in past 10 days. It's done with this simple query: from Book in DB.Books where (from Sale in DB.Sales where Sale.SalesAmount >= 50 && Sale.DateOfSale >= DateTime.Now.AddDays(-10) select Sale.BookID).Contains(Book.ID) select Book The point is, I have to use the checking part in several queries and I decided to create an array with IDs of all popular books: var popularBooksIDs = from Sale in DB.Sales where Sale.SalesAmount >= 50 && Sale.DateOfSale >= DateTime.Now.AddDays(-10) select Sale.BookID; BUT when I try to do the query now: from Book in DB.Books where popularBooksIDs.Contains(Book.ID) select Book It doesn't work! That's why I think that we can't use thins kinds of shortcuts in LINQ to SQL queries, like we can't use them in real SQL. We have to create straightforward queries, am I right?

    Read the article

  • perl dancer: passing database info to template

    - by Bubnoff
    Following Dancer tutorial here: http://search.cpan.org/dist/Dancer/lib/Dancer/Tutorial.pod I'm using my own sqlite3 database with this schema CREATE TABLE if not exists location (location_code TEXT PRIMARY KEY, name TEXT, stations INTEGER); CREATE TABLE if not exists session (id INTEGER PRIMARY KEY, date TEXT, sessions INTEGER, location_code TEXT, FOREIGN KEY(location_code) REFERENCES location(location_code)); My dancer code ( helloWorld.pm ) for the database: package helloWorld; use Dancer; use DBI; use File::Spec; use File::Slurp; use Template; our $VERSION = '0.1'; set 'template' => 'template_toolkit'; set 'logger' => 'console'; my $base_dir = qq(/home/automation/scripts/Area51/perl/dancer); # database crap sub connect_db { my $db = qw(/home/automation/scripts/Area51/perl/dancer/sessions.sqlite); my $dbh = DBI->connect("dbi:SQLite:dbname=$db", "", "", { RaiseError => 1, AutoCommit => 1 }); return $dbh; } sub init_db { my $db = connect_db(); my $file = qq($base_dir/schema.sql); my $schema = read_file($file); $db->do($schema) or die $db->errstr; } get '/' => sub { my $branch_code = qq(BPT); my $dbh = connect_db(); my $sql = q(SELECT * FROM session); my $sth = $dbh->prepare($sql) or die $dbh->errstr; $sth->execute or die $dbh->errstr; my $key_field = q(id); template 'show_entries.tt', { 'branch' => $branch_code, 'data' => $sth->fetchall_hashref($key_field), }; }; init_db(); true; Tried the example template on the site, doesn't work. <% FOREACH id IN data.keys.nsort %> <li>Date is: <% data.$id.sessions %> </li> <% END %> Produces page but with no data. How do I troubleshoot this as no clues come up in the console/cli? Thanks Bubnoff

    Read the article

  • How to define generic super type for static factory method?

    - by Esko
    If this has already been asked, please link and close this one. I'm currently prototyping a design for a simplified API of a certain another API that's a lot more complex (and potentially dangerous) to use. Considering the related somewhat complex object creation I decided to use static factory methods to simplify the API and I currently have the following which works as expected: public class Glue<T> { private List<Type<T>> types; private Glue() { types = new ArrayList<Type<T>>(); } private static class Type<T> { private T value; /* some other properties, omitted for simplicity */ public Type(T value) { this.value = value; } } public static <T> Glue<T> glueFactory(String name, T first, T second) { Glue<T> g = new Glue<T>(); Type<T> firstType = new Glue.Type<T>(first); Type<T> secondType = new Glue.Type<T>(second); g.types.add(firstType); g.types.add(secondType); /* omitted complex stuff */ return g; } } As said, this works as intended. When the API user (=another developer) types Glue<Horse> strongGlue = Glue.glueFactory("2HP", new Horse(), new Horse()); he gets exactly what he wanted. What I'm missing is that how do I enforce that Horse - or whatever is put into the factory method - always implements both Serializable and Comparable? Simply adding them to factory method's signature using <T extends Comparable<T> & Serializable> doesn't necessarily enforce this rule in all cases, only when this simplified API is used. That's why I'd like to add them to the class' definition and then modify the factory method accordingly. PS: No horses (and definitely no ponies!) were harmed in writing of this question.

    Read the article

  • CI pagination, POST problem

    - by Gwood
    Okay, I am pretty new in CI and I am stuck on pagination. I am performing this pagination on a record set that is result of a query. Now everything seems to be working fine. But there’s some problem probably with the link. I am displaying 10 results per page. Now if the results are less than 10 then it’s fine. Or If I pull up the entire records in the table it works fine. But in case the result is more than 10 rows, then the first 10 is perfectly displayed, and when I click on the pagination link to get to the next page the next page displays the rest of the results from the query as well as, other records in the table. ??? I am confused.. Any help?? Here’s the model code I am using .... function getTeesLike($field,$param) { $this-db-like($field,$param); $this-db-limit(10, $this-uri-segment(3)); $query=$this-db-get(‘shirt’); if($query-num_rows()0){ return $query-result_array(); } } function getNumTeesfromQ($field,$param) { $this-db-like($field,$param); $query=$this-db-get(‘shirt’); return $query-num_rows(); } And here’s the controller code .... $KW=$this-input-post(‘searchstr’); $this-load-library(‘pagination’); $config[‘base_url’]=‘http://localhost/cit/index.php/tees/show/’; $config[‘total_rows’]=$this-T-getNumTeesfromQ(‘Title’,$KW); $config[‘per_page’]=‘10’; $this-pagination-initialize($config); $data[‘tees’]=$this-T-getTeesLike(‘Title’,$KW); $data[‘title’]=‘Displaying Tees data’; $data[‘header’]=‘Tees List’; $data[‘links’]=$this-pagination-create_links(); $this-load-view(‘tee_res’, $data); //What am I doing wrong here ???? Pls help ... I guess the problem is with the $KW=$this-input-post(‘searchstr’); .. Because if I hard code a value for $KW it works fine. May be I should use POST differently ..but how do I pass the value from the form without POSTING it , its CI so not GET ... ??????

    Read the article

  • openDatabase Hello World

    - by cf_PhillipSenn
    I'm trying to learn about openDatabase, and I think this I'm getting it to INSERT INTO TABLE1, but I can't verify that the SELECT * FROM TABLE1 is working. <html> <head> <script src="http://www.google.com/jsapi"></script> <script type="text/javascript"> google.load("jquery", "1"); </script> <script type="text/javascript"> var db; $(function(){ db = openDatabase('HelloWorld'); db.transaction( function(transaction) { transaction.executeSql( 'CREATE TABLE IF NOT EXISTS Table1 ' + ' (TableID INTEGER NOT NULL PRIMARY KEY AUTOINCREMENT, ' + ' Field1 TEXT NOT NULL );' ); } ); db.transaction( function(transaction) { transaction.executeSql( 'SELECT * FROM Table1;',function (transaction, result) { for (var i=0; i < result.rows.length; i++) { alert('1'); $('body').append(result.rows.item(i)); } }, errorHandler ); } ); $('form').submit(function() { var xxx = $('#xxx').val(); db.transaction( function(transaction) { transaction.executeSql( 'INSERT INTO Table1 (Field1) VALUES (?);', [xxx], function(){ alert('Saved!'); }, errorHandler ); } ); return false; }); }); function errorHandler(transaction, error) { alert('Oops. Error was '+error.message+' (Code '+error.code+')'); transaction.executeSql('INSERT INTO errors (code, message) VALUES (?, ?);', [error.code, error.message]); return false; } </script> </head> <body> <form method="post"> <input name="xxx" id="xxx" /> <p> <input type="submit" name="OK" /> </p> <a href="http://www.google.com">Cancel</a> </form> </body> </html>

    Read the article

  • Help With LINQ: Mixed Joins and Specifying Default Values

    - by Corey O.
    I am trying to figure out how to do a mixed-join in LINQ with specific access to 2 LINQ objects. Here is an example of how the actual TSQL query might look: SELECT * FROM [User] AS [a] INNER JOIN [GroupUser] AS [b] ON [a].[UserID] = [b].[UserID] INNER JOIN [Group] AS [c] ON [b].[GroupID] = [c].[GroupID] LEFT JOIN [GroupEntries] AS [d] ON [a].[GroupID] = [d].[GroupID] WHERE [a].[UserID] = @UserID At the end, basically what I would like is an enumerable object full of GroupEntry objects. What am interested is the last two tables/objects in this query. I will be displaying Groups as a group header, and all of the Entries underneath their group heading. If there are no entries for a group, I still want to see that group as a header without any entries. Here's what I have so far: So from that I'd like to make a function: public void DisplayEntriesByUser(int user_id) { MyDataContext db = new MyDataContext(); IEnumberable<GroupEntries> entries = ( from user in db.Users where user.UserID == user_id join group_user in db.GroupUsers on user.UserID = group_user.UserID into a from join1 in a join group in db.Groups on join1.GroupID equals group.GroupID into b from join2 in b join entry in db.Entries.DefaultIfEmpty() on join2.GroupID equals entry.GroupID select entry ); Group last_group_id = 0; foreach(GroupEntry entry in entries) { if (last_group_id == 0 || entry.GroupID != last_group_id) { last_group_id = entry.GroupID; System.Console.WriteLine("---{0}---", entry.Group.GroupName.ToString().ToUpper()); } if (entry.EntryID) { System.Console.WriteLine(" {0}: {1}", entry.Title, entry.Text); } } } The example above does not work quite as expected. There are 2 problems that I have not been able to solve: I still seem to be getting an INNER JOIN instead of a LEFT JOIN on the last join. I am not getting any empty results, so groups without entries do not appear. I need to figure out a way so that I can fill in the default values for blank sets of entries. That is, if there is a group without an entry, I would like to have a mostly blank entry returned, except that I'd want the EntryID to be null or 0, the GroupID to be that of of the empty group that it represents, and I'd need a handle on the entry.Group object (i.e. it's parent, empty Group object). Any help on this would be greatly appreciated. Note: Table names and real-world representation were derived purely for this example, but their relations simplify what I'm trying to do.

    Read the article

  • Having trouble doing an Update with a Linq to Sql object

    - by Pure.Krome
    Hi folks, i've got a simple linq to sql object. I grab it from the database and change a field then save. No rows have been updated. :( When I check the full Sql code that is sent over the wire, I notice that it does an update to the row, not via the primary key but on all the fields via the where clause. Is this normal? I would have thought that it would be easy to update the field(s) with the where clause linking on the Primary Key, instead of where'ing (is that a word :P) on each field. here's the code... using (MyDatabase db = new MyDatabase()) { var boardPost = (from bp in db.BoardPosts where bp.BoardPostId == boardPostId select bp).SingleOrDefault(); if (boardPost != null && boardPost.BoardPostId > 0) { boardPost.ListId = listId; // This changes the value from 0 to 'x' db.SubmitChanges(); } } and here's some sample sql.. exec sp_executesql N'UPDATE [dbo].[BoardPost] SET [ListId] = @p6 WHERE ([BoardPostId] = @p0) AND .... <snip the other fields>',N'@p0 int,@p1 int,@p2 nvarchar(9),@p3 nvarchar(10),@p4 int,@p5 datetime,@p6 int',@p0=1276,@p1=212787,@p2=N'ttreterte',@p3=N'ttreterte3',@p4=1,@p5='2009-09-25 12:32:12.7200000',@p6=72 Now, i know there's a datetime field in this update .. and when i checked the DB it's value was/is '2009-09-25 12:32:12.720' (less zero's, than above) .. so i'm not sure if that is messing up the where clause condition... but still! should it do a where clause on the PK's .. if anything .. for speed! Yes / no ? UPDATE After reading nitzmahone's reply, I then tried playing around with the optimistic concurrency on some values, and it still didn't work :( So then I started some new stuff ... with the optimistic concurrency happening, it includes a where clause on the field it's trying to update. When that happens, it doesn't work. so.. in the above sql, the where clause looks like this ... WHERE ([BoardPostId] = @p0) AND ([ListId] IS NULL) AND ... <rest snipped>) This doesn't sound right! the value in the DB is null, before i do the update. but when i add the ListId value to the where clause (or more to the point, when L2S add's it because of the optomistic concurrecy), it fails to find/match the row. wtf?

    Read the article

  • codeIgniter: pass parameter to a select query from previous query

    - by krike
    I'm creating a little management tool for the browser game travian. So I select all the villages from the database and I want to display some content that's unique to each of the villages. But in order to query for those unique details I need to pass the id of the village. How should I do this? this is my code (controller): function members_area() { global $site_title; $this->load->model('membership_model'); if($this->membership_model->get_villages()) { $data['rows'] = $this->membership_model->get_villages(); $id = 1;//this should be dynamic, but how? if($this->membership_model->get_tasks($id)): $data['tasks'] = $this->membership_model->get_tasks($id); endif; } $data['title'] = $site_title." | Your account"; $data['main_content'] = 'account'; $this->load->view('template', $data); } and this is the 2 functions I'm using in the model: function get_villages() { $q = $this->db->get('villages'); if($q->num_rows() > 0) { foreach ($q->result() as $row) { $data[] = $row; } return $data; } } function get_tasks($id) { $this->db->select('name'); $this->db->from('tasks'); $this->db->where('villageid', $id); $q = $this->db->get(); if($q->num_rows() > 0) { foreach ($q->result() as $task) { $data[] = $task; } return $data; } } and of course the view: <?php foreach($rows as $r) : ?> <div class="village"> <h3><?php echo $r->name; ?></h3> <ul> <?php foreach($tasks as $task): ?> <li><?php echo $task->name; ?></li> <?php endforeach; ?> </ul> <?php echo anchor('site/add_village/'.$r->id.'', '+ add new task'); ?> </div> <?php endforeach; ?> ps: please do not remove the comment in the first block of code!

    Read the article

  • MySQL Database Query - Codeigniter

    - by user2450349
    I am building an application with Codeigniter and need some help with a DB query. I have a table called users with the following fields: user_id, user_name, user_password, user_email, user_role, user_manager_id In my app, I pull all records from the user table using the following: function get_clients() { $this->db->select('*'); $this->db->where('user_role', 'client'); $this->db->order_by("user_name", "Asc"); $query = $this->db->get("users"); return $query->result_array(); } This works as expected, however when I display the results in the view, I also want to display a new column called Manager which will display the managers user_name field. The user_manager_id is the id of the user from the same table. Im guessing you can create an outer join on the same table but not sure. In the view, I am displaying the returned info as follows: <table class="table table-striped" id="zero-configuration"> <thead> <tr> <th>Name</th> <th>Email</th> <th>Manager</th> </tr> </thead> <tbody> <?php foreach($clients as $row) { ?> <tr> <td><?php echo $row['user_name']; ?> (<?php echo $row['user_username']; ?>)</td> <td><?php echo $row['user_email']; ?></td> <td><?php echo $row['???']; ?></td> </tr> <?php } ?> </tbody> </table> Any idea of how I can form the query and display the manager name is the view? Example: user_id user_name user_password user_email user_role user_manager_id 1 Ollie adjjk34jcd [email protected] client null 2 James djklsdfsdjk [email protected] client 1 When i query the database, i want to display results like this: Ollie [email protected] James [email protected] Ollie

    Read the article

  • Delete duplicate rows, do not preserve one row

    - by Radley
    I need a query that goes through each entry in a database, checks if a single value is duplicated elsewhere in the database, and if it is - deletes both entries (or all, if more than two). Problem is the entries are URLs, up to 255 characters, with no way of identifying the row. Some existing answers on Stackoverflow do not work for me due to performance limitations, or they use uniqueid which obviously won't work when dealing with a string. Long Version: I have two databases containing URLs (and only URLs). One database has around 3,000 urls and the other around 1,000. However, a large majority of the 1,000 urls were taken from the 3,000 url database. I need to merge the 1,000 into the 3,000 as new entries only. For this, I made a third database with combined URLs from both tables, about 4,000 entries. I need to find all duplicate entries in this database and delete them (Both of them, without leaving either). I have followed the query of a few examples on this site, but whenever I try to delete both entries it ends up deleting all the entries, or giving sql errors. Alternatively: I have two databases, each containing the separate database. I need to check each row from one database against the other to find any that aren't duplicates, and then add those to a third database. Edit: I've got my own PHP solution which is pretty hacky, but works. I cannot answer my own question for 8 hours because I'm new, so here it is for now: I went with a PHP script to accomplish this, as I'm more familiar with PHP than MySQL. This generates a simple list of urls that only exist in the target database, but not both. If you have more than 7,000 entries to parse this may take awhile, and you will need to copy/paste the results into a text file or expand the script to store them back into a database. I'm just doing it manually to save time. Note: Uses MeekroDB <pre> <?php require('meekrodb.2.1.class.php'); DB::$user = 'root'; DB::$password = ''; DB::$dbName = 'testdb'; $all = DB::query('SELECT * FROM old_urls LIMIT 7000'); foreach($all as $row) { $test = DB::query('SELECT url FROM new_urls WHERE url=%s', $row['url']); if (!is_array($test)) { echo $row['url'] . "\n"; }else{ if (count($test) == 0) { echo $row['url'] . "\n"; } } } ?> </pre>

    Read the article

  • PHP-Mcrypt Installation

    - by Infinity
    I need to install php-mcrypt on my CentOS 5.5 VPS, When I try yum install php-mcrypt, it says that it is set to be updated which implies that it is already installed. I looked in the /usr/lib/php/modules and cant find the .so file. Anyway I want to update it but yum is giving the following error, I am running PHP-FPM on Nginx. Last login: Thu Apr 21 12:13:30 2011 from cpc2-seve18-2-0-cust438.13-3.cable.virginmedia.com [root@infinity ~]# yum install php-mcrypt Setting up Install Process Resolving Dependencies --> Running transaction check ---> Package php-mcrypt.i386 0:5.1.6-15.el5.centos.1 set to be updated --> Processing Dependency: php-api = 20041225 for package: php-mcrypt --> Processing Dependency: php >= 5.1.6 for package: php-mcrypt --> Running transaction check ---> Package php.i386 0:5.1.6-27.el5_5.3 set to be updated --> Processing Dependency: php-common = 5.1.6-27.el5_5.3 for package: php --> Processing Dependency: php-cli = 5.1.6-27.el5_5.3 for package: php ---> Package php-mcrypt.i386 0:5.1.6-15.el5.centos.1 set to be updated --> Processing Dependency: php-api = 20041225 for package: php-mcrypt --> Running transaction check ---> Package php.i386 0:5.1.6-27.el5_5.3 set to be updated --> Processing Dependency: php-common = 5.1.6-27.el5_5.3 for package: php ---> Package php-cli.i386 0:5.1.6-27.el5_5.3 set to be updated --> Processing Dependency: php-common = 5.1.6-27.el5_5.3 for package: php-cli ---> Package php-mcrypt.i386 0:5.1.6-15.el5.centos.1 set to be updated --> Processing Dependency: php-api = 20041225 for package: php-mcrypt --> Finished Dependency Resolution php-mcrypt-5.1.6-15.el5.centos.1.i386 from extras has depsolving problems --> Missing Dependency: php-api = 20041225 is needed by package php-mcrypt-5.1.6-15.el5.centos.1.i386 (extras) php-5.1.6-27.el5_5.3.i386 from base has depsolving problems --> Missing Dependency: php-common = 5.1.6-27.el5_5.3 is needed by package php-5.1.6-27.el5_5.3.i386 (base) php-cli-5.1.6-27.el5_5.3.i386 from base has depsolving problems --> Missing Dependency: php-common = 5.1.6-27.el5_5.3 is needed by package php-cli-5.1.6-27.el5_5.3.i386 (base) Error: Missing Dependency: php-api = 20041225 is needed by package php-mcrypt-5.1.6-15.el5.centos.1.i386 (extras) Error: Missing Dependency: php-common = 5.1.6-27.el5_5.3 is needed by package php-cli-5.1.6-27.el5_5.3.i386 (base) Error: Missing Dependency: php-common = 5.1.6-27.el5_5.3 is needed by package php-5.1.6-27.el5_5.3.i386 (base) You could try using --skip-broken to work around the problem You could try running: package-cleanup --problems package-cleanup --dupes rpm -Va --nofiles --nodigest The program package-cleanup is found in the yum-utils package. [root@infinity ~]# Any ideas?

    Read the article

  • dns queries not using nscd for caching

    - by xenoterracide
    I'm trying to use nscd (Nameservices Cache Daemon) to cache dns locally so I can stop using bind to do it. I've gotten it started and ntpd seems to attempt to use it. But everything else for hosts seems to ignore it. e.g if I do dig apache.org 3 times none of them will hit the cache. I'm viewing the cache stats using nscd -g to determine whether it's been used. I've also turned the debug log level up to see if I can see it hitting and the queries don't even hit nscd. nsswitch.conf # Begin /etc/nsswitch.conf passwd: files group: files shadow: files publickey: files hosts: cache files dns networks: files protocols: files services: files ethers: files rpc: files netgroup: files # End /etc/nsswitch.confenter code here nscd.conf # # /etc/nscd.conf # # An example Name Service Cache config file. This file is needed by nscd. # # Legal entries are: # # logfile <file> # debug-level <level> # threads <initial #threads to use> # max-threads <maximum #threads to use> # server-user <user to run server as instead of root> # server-user is ignored if nscd is started with -S parameters # stat-user <user who is allowed to request statistics> # reload-count unlimited|<number> # paranoia <yes|no> # restart-interval <time in seconds> # # enable-cache <service> <yes|no> # positive-time-to-live <service> <time in seconds> # negative-time-to-live <service> <time in seconds> # suggested-size <service> <prime number> # check-files <service> <yes|no> # persistent <service> <yes|no> # shared <service> <yes|no> # max-db-size <service> <number bytes> # auto-propagate <service> <yes|no> # # Currently supported cache names (services): passwd, group, hosts, services # logfile /var/log/nscd.log threads 4 max-threads 32 server-user nobody # stat-user somebody debug-level 9 # reload-count 5 paranoia no # restart-interval 3600 enable-cache passwd yes positive-time-to-live passwd 600 negative-time-to-live passwd 20 suggested-size passwd 211 check-files passwd yes persistent passwd yes shared passwd yes max-db-size passwd 33554432 auto-propagate passwd yes enable-cache group yes positive-time-to-live group 3600 negative-time-to-live group 60 suggested-size group 211 check-files group yes persistent group yes shared group yes max-db-size group 33554432 auto-propagate group yes enable-cache hosts yes positive-time-to-live hosts 3600 negative-time-to-live hosts 20 suggested-size hosts 211 check-files hosts yes persistent hosts yes shared hosts yes max-db-size hosts 33554432 enable-cache services yes positive-time-to-live services 28800 negative-time-to-live services 20 suggested-size services 211 check-files services yes persistent services yes shared services yes max-db-size services 33554432 resolv.conf # Generated by dhcpcd from eth0 nameserver 127.0.0.1 domain westell.com nameserver 192.168.1.1 nameserver 208.67.222.222 nameserver 208.67.220.220 as kind of a side note I'm using archlinux.

    Read the article

  • dns queries not using nscd for caching

    - by xenoterracide
    I'm trying to use nscd (Nameservices Cache Daemon) to cache dns locally so I can stop using bind to do it. I've gotten it started and ntpd seems to attempt to use it. But everything else for hosts seems to ignore it. e.g if I do dig apache.org 3 times none of them will hit the cache. I'm viewing the cache stats using nscd -g to determine whether it's been used. I've also turned the debug log level up to see if I can see it hitting and the queries don't even hit nscd. nsswitch.conf # Begin /etc/nsswitch.conf passwd: files group: files shadow: files publickey: files hosts: cache files dns networks: files protocols: files services: files ethers: files rpc: files netgroup: files # End /etc/nsswitch.confenter code here nscd.conf # # /etc/nscd.conf # # An example Name Service Cache config file. This file is needed by nscd. # # Legal entries are: # # logfile <file> # debug-level <level> # threads <initial #threads to use> # max-threads <maximum #threads to use> # server-user <user to run server as instead of root> # server-user is ignored if nscd is started with -S parameters # stat-user <user who is allowed to request statistics> # reload-count unlimited|<number> # paranoia <yes|no> # restart-interval <time in seconds> # # enable-cache <service> <yes|no> # positive-time-to-live <service> <time in seconds> # negative-time-to-live <service> <time in seconds> # suggested-size <service> <prime number> # check-files <service> <yes|no> # persistent <service> <yes|no> # shared <service> <yes|no> # max-db-size <service> <number bytes> # auto-propagate <service> <yes|no> # # Currently supported cache names (services): passwd, group, hosts, services # logfile /var/log/nscd.log threads 4 max-threads 32 server-user nobody # stat-user somebody debug-level 9 # reload-count 5 paranoia no # restart-interval 3600 enable-cache passwd yes positive-time-to-live passwd 600 negative-time-to-live passwd 20 suggested-size passwd 211 check-files passwd yes persistent passwd yes shared passwd yes max-db-size passwd 33554432 auto-propagate passwd yes enable-cache group yes positive-time-to-live group 3600 negative-time-to-live group 60 suggested-size group 211 check-files group yes persistent group yes shared group yes max-db-size group 33554432 auto-propagate group yes enable-cache hosts yes positive-time-to-live hosts 3600 negative-time-to-live hosts 20 suggested-size hosts 211 check-files hosts yes persistent hosts yes shared hosts yes max-db-size hosts 33554432 enable-cache services yes positive-time-to-live services 28800 negative-time-to-live services 20 suggested-size services 211 check-files services yes persistent services yes shared services yes max-db-size services 33554432 resolv.conf # Generated by dhcpcd from eth0 nameserver 127.0.0.1 domain westell.com nameserver 192.168.1.1 nameserver 208.67.222.222 nameserver 208.67.220.220 as kind of a side note I'm using archlinux.

    Read the article

  • Trying to connect phpMyAdmin to remote mySQL server ( 2002: can't connect )

    - by Malcolm Jones
    Trying to get phpMyAdmin to talk to a remote mySQL server. The config is below and there is already a user set up in mySQL DB to be able to log in from the specified host that PMA sits on. Hosting is provided by Rackspace (Rightscale) and both cloud servers behind the same firewall. [config.inc.php] <?php $cfg['blowfish_secret'] = ''; $i = 0; $i++; $cfg['Servers'][$i]['host'] = 'XX.XX.XX.XX'; // MySQL hostname or IP address $cfg['Servers'][$i]['port'] = ''; // MySQL port - leave blank for default port $cfg['Servers'][$i]['socket'] = ''; // Path to the socket - leave blank for default socket $cfg['Servers'][$i]['connect_type'] = 'tcp'; // How to connect to MySQL server ('tcp' or 'socket') $cfg['Servers'][$i]['extension'] = 'mysql'; // The php MySQL extension to use ('mysql' or 'mysqli') $cfg['Servers'][$i]['compress'] = FALSE; // Use compressed protocol for the MySQL connection // (requires PHP >= 4.3.0) $cfg['Servers'][$i]['controluser'] = ''; // MySQL control user settings // (this user must have read-only $cfg['Servers'][$i]['controlpass'] = ''; // access to the "mysql/user" // and "mysql/db" tables). // The controluser is also // used for all relational // features (pmadb) $cfg['Servers'][$i]['auth_type'] = 'config'; // Authentication method (config, http or cookie based)? $cfg['Servers'][$i]['user'] = 'USERNAME'; // MySQL user $cfg['Servers'][$i]['password'] = 'PASSWORD'; // MySQL password (only needed // with 'config' auth_type) $cfg['Servers'][$i]['only_db'] = ''; // If set to a db-name, only // this db is displayed in left frame // It may also be an array of db-names, where sorting order is relevant. $cfg['Servers'][$i]['hide_db'] = ''; // Database name to be hidden from listings $cfg['Servers'][$i]['verbose'] = ''; // Verbose name for this host - leave blank to show the hostname $cfg['Servers'][$i]['pmadb'] = ''; // Database used for Relation, Bookmark and PDF Features // (see scripts/create_tables.sql) // - leave blank for no support // DEFAULT: 'phpmyadmin' $cfg['Servers'][$i]['bookmarktable'] = ''; // Bookmark table // - leave blank for no bookmark support // DEFAULT: 'pma_bookmark' $cfg['Servers'][$i]['relation'] = ''; // table to describe the relation between links (see doc) // - leave blank for no relation-links support // DEFAULT: 'pma_relation' $cfg['Servers'][$i]['table_info'] = ''; // table to describe the display fields // - leave blank for no display fields support // DEFAULT: 'pma_table_info' $cfg['Servers'][$i]['table_coords'] = ''; // table to describe the tables position for the PDF schema // - leave blank for no PDF schema support // DEFAULT: 'pma_table_coords' $cfg['Servers'][$i]['pdf_pages'] = ''; // table to describe pages of relationpdf // - leave blank if you don't want to use this // DEFAULT: 'pma_pdf_pages' $cfg['Servers'][$i]['column_info'] = ''; // table to store column information // - leave blank for no column comments/mime types // DEFAULT: 'pma_column_info' $cfg['Servers'][$i]['history'] = ''; // table to store SQL history // - leave blank for no SQL query history // DEFAULT: 'pma_history' $cfg['Servers'][$i]['verbose_check'] = TRUE; // set to FALSE if you know that your pma_* tables // are up to date. This prevents compatibility // checks and thereby increases performance. $cfg['Servers'][$i]['AllowRoot'] = TRUE; // whether to allow root login $cfg['Servers'][$i]['AllowDeny']['order'] // Host authentication order, leave blank to not use = ''; $cfg['Servers'][$i]['AllowDeny']['rules'] // Host authentication rules, leave blank for defaults = array(); Please let me know if you need anymore info. -- Malcolm

    Read the article

  • GPGPU

    WhatGPU obviously stands for Graphics Processing Unit (the silicon powering the display you are using to read this blog post). The extra GP in front of that stands for General Purpose computing.So, altogether GPGPU refers to computing we can perform on GPU for purposes beyond just drawing on the screen. In effect, we can use a GPGPU a bit like we already use a CPU: to perform some calculation (that doesn’t have to have any visual element to it). The attraction is that a GPGPU can be orders of magnitude faster than a CPU.WhyWhen I was at the SuperComputing conference in Portland last November, GPGPUs were all the rage. A quick online search reveals many articles introducing the GPGPU topic. I'll just share 3 here: pcper (ignoring all pages except the first, it is a good consumer perspective), gizmodo (nice take using mostly layman terms) and vizworld (answering the question on "what's the big deal").The GPGPU programming paradigm (from a high level) is simple: in your CPU program you define functions (aka kernels) that take some input, can perform the costly operation and return the output. The kernels are the things that execute on the GPGPU leveraging its power (and hence execute faster than what they could on the CPU) while the host CPU program waits for the results or asynchronously performs other tasks.However, GPGPUs have different characteristics to CPUs which means they are suitable only for certain classes of problem (i.e. data parallel algorithms) and not for others (e.g. algorithms with branching or recursion or other complex flow control). You also pay a high cost for transferring the input data from the CPU to the GPU (and vice versa the results back to the CPU), so the computation itself has to be long enough to justify the overhead transfer costs. If your problem space fits the criteria then you probably want to check out this technology.HowSo where can you get a graphics card to start playing with all this? At the time of writing, the two main vendors ATI (owned by AMD) and NVIDIA are the obvious players in this industry. You can read about GPGPU on this AMD page and also on this NVIDIA page. NVIDIA's website also has a free chapter on the topic from the "GPU Gems" book: A Toolkit for Computation on GPUs.If you followed the links above, then you've already come across some of the choices of programming models that are available today. Essentially, AMD is offering their ATI Stream technology accessible via a language they call Brook+; NVIDIA offers their CUDA platform which is accessible from CUDA C. Choosing either of those locks you into the GPU vendor and hence your code cannot run on systems with cards from the other vendor (e.g. imagine if your CPU code would run on Intel chips but not AMD chips). Having said that, both vendors plan to support a new emerging standard called OpenCL, which theoretically means your kernels can execute on any GPU that supports it. To learn more about all of these there is a website: gpgpu.org. The caveat about that site is that (currently) it completely ignores the Microsoft approach, which I touch on next.On Windows, there is already a cross-GPU-vendor way of programming GPUs and that is the DirectX API. Specifically, on Windows Vista and Windows 7, the DirectX 11 API offers a dedicated subset of the API for GPGPU programming: DirectCompute. You use this API on the CPU side, to set up and execute the kernels that run on the GPU. The kernels are written in a language called HLSL (High Level Shader Language). You can use DirectCompute with HLSL to write a "compute shader", which is the term DirectX uses for what I've been referring to in this post as a "kernel". For a comprehensive collection of links about this (including tutorials, videos and samples) please see my blog post: DirectCompute.Note that there are many efforts to build even higher level languages on top of DirectX that aim to expose GPGPU programming to a wider audience by making it as easy as today's mainstream programming models. I'll mention here just two of those efforts: Accelerator from MSR and Brahma by Ananth. Comments about this post welcome at the original blog.

    Read the article

  • What I don&rsquo;t like about WIF&rsquo;s Claims-based Authorization

    - by Your DisplayName here!
    In my last post I wrote about what I like about WIF’s proposed approach to authorization – I also said that I definitely would build upon that infrastructure for my own systems. But implementing such a system is a little harder as it could be. Here’s why (and that’s purely my perspective): First of all WIF’s authorization comes in two “modes” Per-request authorization. When an ASP.NET/WCF request comes in, the registered authorization manager gets called. For SOAP the SOAP action gets passed in. For HTTP requests (ASP.NET, WCF REST) the URL and verb. Imperative authorization This happens when you explicitly call the claims authorization API from within your code. There you have full control over the values for action and resource. In ASP.NET per-request authorization is optional (depends on if you have added the ClaimsAuthorizationHttpModule). In WCF you always get the per-request checks as soon as you register the authorization manager in configuration. I personally prefer the imperative authorization because first of all I don’t believe in URL based authorization. Especially in the times of MVC and routing tables, URLs can be easily changed – but then you also have to adjust your authorization logic every time. Also – you typically need more knowledge than a simple “if user x is allowed to invoke operation x”. One problem I have is, both the per-request calls as well as the standard WIF imperative authorization APIs wrap actions and resources in the same claim type. This makes it hard to distinguish between the two authorization modes in your authorization manager. But you typically need that feature to structure your authorization policy evaluation in a clean way. The second problem (which is somehow related to the first one) is the standard API for interacting with the claims authorization manager. The API comes as an attribute (ClaimsPrincipalPermissionAttribute) as well as a class to use programmatically (ClaimsPrincipalPermission). Both only allow to pass in simple strings (which results in the wrapping with standard claim types mentioned earlier). Both throw a SecurityException when the check fails. The attribute is a code access permission attribute (like PrincipalPermission). That means it will always be invoked regardless how you call the code. This may be exactly what you want, or not. In a unit testing situation (like an MVC controller) you typically want to test the logic in the function – not the security check. The good news is, the WIF API is flexible enough that you can build your own infrastructure around their core. For my own projects I implemented the following extensions: A way to invoke the registered claims authorization manager with more overloads, e.g. with different claim types or a complete AuthorizationContext. A new CAS attribute (with the same calling semantics as the built-in one) with custom claim types. A MVC authorization attribute with custom claim types. A way to use branching – as opposed to catching a SecurityException. I will post the code for these various extensions here – so stay tuned.

    Read the article

< Previous Page | 255 256 257 258 259 260 261 262 263 264 265 266  | Next Page >