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• Problem getting ar_mailer/ar_sendmail working on new server

  - by Max Williams

  Hey all. I've got a new app up and running on a new ubuntu server. It's working fine generally but i can't get ar_sendmail working. I'm following the instructions on this page: http://www.ameravant.com/posts/sending-tons-of-emails-in-ruby-on-rails-with-ar_mailer The setup is all done, ie i can "deliver mails" which just saves records in my Email table. Now i want to get the ar_sendmail daemon running to actually send them. (so i'm at 'Running ar_sendmail in daemon mode' in that web page). First thing: ar_sendmail --mailq >>ar_sendmail: command not found Ok...so, where is ar_sendmail? I have a look and there's an ar_sendmail file in the bin folder of the ar_mailer plugin, so i add the location of that to my path. I don't know if this was the right thing to do or not. Ok, so try again. ar_sendmail --mailq /var/www/apps/millionaire/vendor/plugins/ar_mailer/bin/ar_sendmail:3:in `require': no such file to load -- action_mailer/ar_sendmail (LoadError) from /var/www/apps/millionaire/vendor/plugins/ar_mailer/bin/ar_sendmail:3 hmm. Here's the offending file, there's not much there. #!/usr/bin/env ruby require 'action_mailer/ar_sendmail' ActionMailer::ARSendmail.run ok...so it literally is just trying to require this and can't find it. The file, action_mailer/ar_sendmail.rb is in the ar_mailer plugin, in it's lib folder. So, given that it's being called from inside the plugin, it should be able to see this right? I've got a feeling that i'm way off the track here and have missed something simple. Can anyone set me straight? I'm using rails 2.3.4 in case that's relevant. EDIT - i just realised something kind of dumb: when i call ar_sendmail from the command line like this, i'm just loading that one file, which doesn't know where it's supposed to look for the rest of the stuff, i think. Which really makes me think that i'm not trying to run the right thing. Is the ar_sendmail daemon a seperate program altogether, that i would get with apt_get or something? EDIT2 - i made some progress by installing the ar_mailer gem (which the guide said i shouldn't do) and that does seem to run. It's sending some mail request somewhere and clearing the Email table of pending emails. Running ar_sendmail in -ov (oneshot verbal) mode i see it report this for example: sent email 00000000019 from [email protected] to [email protected]: # So, it actually looks like it's working now and i just need to set up the ACTUAL THING WHICH SENDS EMAILS. sigh. still grateful for any advice. thanks, max

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• Potential issues using member's "from" address and the "sender" header

  - by Paul Burney

  Hi all, A major component of our application sends email to members on behalf of other members. Currently we set the "From" address to our system address and use a "Reply-to" header with the member's address. The issue is that replies from some email clients (and auto-replies/bounces) don't respect the "Reply-to" header so get sent to our system address, effectively sending them to a black hole. We're considering setting the "From" address to our member's address, and the "Sender" address to our system address. It appears this way would pass SPF and Sender-ID checks. Are there any reasons not to switch to this method? Are there any other potential issues? Thanks in advance, -Paul Here are way more details than you probably need: When the application was first developed, we just changed the "from" address to be that of the sending member as that was the common practice at the time (this was many years ago). We later changed that to have the "from" address be the member's name and our address, i.e., From: "Mary Smith" <[email protected]> With a "reply-to" header set to the member's address: Reply-To: "Mary Smith" <[email protected]> This helped with messages being mis-categorized as spam. As SPF became more popular, we added an additional header that would work in conjunction with our SPF records: Sender: <[email protected]> Things work OK, but it turns out that, in practice, some email clients and most MTA's don't respect the "Reply-To" header. Because of this, many members send messages to [email protected] instead of the desired member. So, I started envisioning various schemes to add data about the sender to the email headers or encode it in the "from" email address so that we could process the response and redirect appropriately. For example, From: "Mary Smith" <[email protected]> where the string after "messages" is a hash representing Mary Smith's member in our system. Of course, that path could lead to a lot of pain as we need to develop MTA functionality for our system address. I was looking again at the SPF documentation and found this page interesting: http://www.openspf.org/Best_Practices/Webgenerated They show two examples, that of evite.com and that of egreetings.com. Basically, evite.com is doing it the way we're doing it. The egreetings.com example uses the member's from address with an added "Sender" header. So the question is, are there any potential issues with using the egreetings method of the member's from address with a sender header? That would eliminate the replies that bad clients send to the system address. I don't believe that it solves the bounce/vacation/whitelist issue since those often send to the MAIL FROM even if Return Path is specified.

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• H.264 over RTP - Identify SPS and PPS Frames

  - by Toby

  I have a raw H.264 Stream from an IP Camera packed in RTP frames. I want to get raw H.264 data into a file so I can convert it with ffmpeg. So when I want to write the data into my raw H.264 file I found out it has to look like this: 00 00 01 [SPS] 00 00 01 [PPS] 00 00 01 [NALByte] [PAYLOAD RTP Frame 1] // Payload always without the first 2 Bytes -> NAL [PAYLOAD RTP Frame 2] [... until PAYLOAD Frame with Mark Bit received] // From here its a new Video Frame 00 00 01 [NAL BYTE] [PAYLOAD RTP Frame 1] .... So I get the SPS and the PPS from the Session Description Protocol out of my preceding RTSP communication. Additionally the camera sends the SPS and the PPSin two single messages before starting with the video stream itself. So I capture the messages in this order: 1. Preceding RTSP Communication here ( including SDP with SPS and PPS ) 2. RTP Frame with Payload: 67 42 80 28 DA 01 40 16 C4 // This is the SPS 3. RTP Frame with Payload: 68 CE 3C 80 // This is the PPS 4. RTP Frame with Payload: ... // Video Data Then there come some Frames with Payload and at some point a RTP Frame with the Marker Bit = 1. This means ( if I got it right) that I have a complete video frame. Afer this I write the Prefix Sequence ( 00 00 01 ) and the NALfrom the payload again and go on with the same procedure. Now my camera sends me after every 8 complete Video Frames the SPS and the PPS again. ( Again in two RTP Frames, as seen in the example above ). I know that especially the PPS can change in between streaming but that's not the problem. My questions are now: 1. Do I need to write the SPS/PPS every 8th Video Frame? If my SPS and my PPS don't change it should be enough to have them written at the very beginning of my file and nothing more? 2. How to distinguish between SPS/PPS and normal RTP Frames? In my C++ Code which parses the transmitted data I need make a difference between the RTP Frames with normal Payload an the ones carrying the SPS/PPS. How can I distinguish them? Okay the SPS/PPS frames are usually way smaller, but that's not a save call to rely on. Because if I ignore them I need to know which data I can throw away, or if I need to write them I need to put the 00 00 01 Prefix in front of them. ? Or is it a fixed rule that they occur every 8th Video Frame?

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• C# Finalize/Dispose pattern

  - by robUK

  Hello, C# 2008 I have been working on this for a while now. And I am still confused about some issues. My questions below 1) I know that you only need a finalizer if you are disposing of unmanaged resources. However, if you are using managed resources that make calls to unmanaged resources. Would you still need to implement a finalizer? 2) However, if you develop a class that doesn't use any unmanged resources directly or indirectly and you implement the IDisposable so that clients of your class can use the 'using statement'. Would it be acceptable to implement the IDisposable just so that clients of your class can use the using statement? using(myClass objClass = new myClass()) { // Do stuff here } 3) I have developed this simple code below to demostrate the Finalize/dispose pattern: public class NoGateway : IDisposable { private WebClient wc = null; public NoGateway() { wc = new WebClient(); wc.DownloadStringCompleted += wc_DownloadStringCompleted; } // Start the Async call to find if NoGateway is true or false public void NoGatewayStatus() { // Start the Async's download // Do other work here wc.DownloadStringAsync(new Uri(www.xxxx.xxx)); } private void wc_DownloadStringCompleted(object sender, DownloadStringCompletedEventArgs e) { // Do work here } // Dispose of the NoGateway object public void Dispose() { wc.DownloadStringCompleted -= wc_DownloadStringCompleted; wc.Dispose(); GC.SuppressFinalize(this); } } Question about the source code: 1) Here I have not added the finalizer. And normally the finalizer will be called by the GC, and the finalizer will call the Dispose. As I don't have the finalizer, when do I call the Dispose method? Is it the client of the class that has to call it? So my class in the example is called NoGateway and the client could use and dispose of the class like this: Would the Dispose method be automatically called when execution reaches the end of the using block? using(NoGateway objNoGateway = new NoGateway()) { // Do stuff here } Or does the client have to manually call the dispose method i.e.? NoGateway objNoGateway = new NoGateway(); // Do stuff with object objNoGateway.Dispose(); // finished with it Many thanks for helping with all these questions, 2) I am using the webclient class in my 'NoGateway' class. Because the webclient implements the IDisposable interface. Does this mean that the webclient indirectly uses unmanaged resources? Is there any hard and fast rule to follow about this. How do I know that a class uses unmanaged resources?

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• Multiple views and source list in a Core Data app

  - by Ellie P.

  I'm working on my first major Cocoa app for an undergraduate research project. The application is document-based and uses Core Data. One of the entities is an abstract entity, Page. Page is parent of several types of pages: ie PageWithHeaderAndFooter, PageWithTwoColumns, BasicPage etc. Page has attributes, such as title and author, that all pages have in common. Each specific type of page has a certain number of layout blocks (PageWithHeaderAndFooter has three: header, footer, body. BasicPage has one: body. etc.) Additionally, all Page subclasses define layout-specific implementations of certain methods. The other relevant entity is Style, which defines the visual look of a Page. (Think of Pages as HTML and Style as CSS.) I would like my app to have an iTunes/Mail-like source list with sections. (One section would be Pages, the other would be Styles.) I have a pretty good idea how to do the sectioned source list (this was a great help). However, after hours of headbanging and fruitless googling, here's what I can't figure out: Pages and Styles listed in the source list, and when you select one of them, all of the relevant fields for that object appear at the right (mostly NSTextViews, pop up menus, etc). I laid that out and did all of the bindings in Interface Builder. The problem is, if my source list contains different types of pages, how do I get a different view to display at the right depending on the type of page selected? For example, if a BasicPage is selected, I want just what you see above: the general page stuff and one NSTextView that corresponds to the one field body of BasicPage. But if I select a PageWithHeaderAndFooter, I want to display the general page stuff plus three NSTextViews (one for header, body, and footer.) If I have a Style selected, I want to display various pop up menus, color wells, etc. For the pages at least, we're only talking about one or more NSTextViews, each of which corresponds to a String attribute of the respective entity. How would you do this? Thank you for your help!

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• Asynchronous vs Synchronous vs Threading in an iPhone App

  - by Coocoo4Cocoa

  I'm in the design stage for an app which will utilize a REST web service and sort of have a dilemma in as far as using asynchronous vs synchronous vs threading. Here's the scenario. Say you have three options to drill down into, each one having its own REST-based resource. I can either lazily load each one with a synchronous request, but that'll block the UI and prevent the user from hitting a back navigation button while data is retrieved. This case applies almost anywhere except for when your application requires a login screen. I can't see any reason to use synchronous HTTP requests vs asynchronous because of that reason alone. The only time it makes sense is to have a worker thread make your synchronous request, and notify the main thread when the request is done. This will prevent the block. The question then is bench marking your code and seeing which has more overhead, a threaded synchronous request or an asynchronous request. The problem with asynchronous requests is you need to either setup a smart notification or delegate system as you can have multiple requests for multiple resources happening at any given time. The other problem with them is if I have a class, say a singleton which is handling all of my data, I can't use asynchronous requests in a getter method. Meaning the following won't go: - (NSArray *)users { if(users == nil) users = do_async_request // NO GOOD return users; } whereas the following: - (NSArray *)users { if(users == nil) users == do_sync_request // OK. return users; } You also might have priority. What I mean by priority is if you look at Apple's Mail application on the iPhone, you'll notice they first suck down your entire POP/IMAP tree before making a second request to retrieve the first 2 lines (the default) of your message. I suppose my question to you experts is this. When are you using asynchronous, synchronous, threads -- and when are you using either async/sync in a thread? What kind of delegation system do you have setup to know what to do when a async request completes? Are you prioritizing your async requests? There's a gamut of solutions to this all too common problem. It's simple to hack something out. The problem is, I don't want to hack and I want to have something that's simple and easy to maintain.

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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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• agent-based simulation: performance issue: Python vs NetLogo & Repast

  - by max

  I'm replicating a small piece of Sugarscape agent simulation model in Python 3. I found the performance of my code is ~3 times slower than that of NetLogo. Is it likely the problem with my code, or can it be the inherent limitation of Python? Obviously, this is just a fragment of the code, but that's where Python spends two-thirds of the run-time. I hope if I wrote something really inefficient it might show up in this fragment: UP = (0, -1) RIGHT = (1, 0) DOWN = (0, 1) LEFT = (-1, 0) all_directions = [UP, DOWN, RIGHT, LEFT] # point is just a tuple (x, y) def look_around(self): max_sugar_point = self.point max_sugar = self.world.sugar_map[self.point].level min_range = 0 random.shuffle(self.all_directions) for r in range(1, self.vision+1): for d in self.all_directions: p = ((self.point[0] + r * d[0]) % self.world.surface.length, (self.point[1] + r * d[1]) % self.world.surface.height) if self.world.occupied(p): # checks if p is in a lookup table (dict) continue if self.world.sugar_map[p].level > max_sugar: max_sugar = self.world.sugar_map[p].level max_sugar_point = p if max_sugar_point is not self.point: self.move(max_sugar_point) Roughly equivalent code in NetLogo (this fragment does a bit more than the Python function above): ; -- The SugarScape growth and motion procedures. -- to M ; Motion rule (page 25) locals [ps p v d] set ps (patches at-points neighborhood) with [count turtles-here = 0] if (count ps > 0) [ set v psugar-of max-one-of ps [psugar] ; v is max sugar w/in vision set ps ps with [psugar = v] ; ps is legal sites w/ v sugar set d distance min-one-of ps [distance myself] ; d is min dist from me to ps agents set p random-one-of ps with [distance myself = d] ; p is one of the min dist patches if (psugar >= v and includeMyPatch?) [set p patch-here] setxy pxcor-of p pycor-of p ; jump to p set sugar sugar + psugar-of p ; consume its sugar ask p [setpsugar 0] ; .. setting its sugar to 0 ] set sugar sugar - metabolism ; eat sugar (metabolism) set age age + 1 end On my computer, the Python code takes 15.5 sec to run 1000 steps; on the same laptop, the NetLogo simulation running in Java inside the browser finishes 1000 steps in less than 6 sec. EDIT: Just checked Repast, using Java implementation. And it's also about the same as NetLogo at 5.4 sec. Recent comparisons between Java and Python suggest no advantage to Java, so I guess it's just my code that's to blame? EDIT: I understand MASON is supposed to be even faster than Repast, and yet it still runs Java in the end.

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• Data access pattern, combining push and pull?

  - by andlju

  I need some advice on what kind of pattern(s) I should use for pushing/pulling data into my application. I'm writing a rule-engine that needs to hold quite a large amount of data in-memory in order to be efficient enough. I have some rather conflicting requirements; It is not acceptable for the engine to always have to wait for a full pre-load of all data before it is functional. Only fetching and caching data on-demand will lead to the engine taking too long before it is running quickly enough. An external event can trigger the need for specific parts of the data to be reloaded. Basically, I think I need a combination of pushing and pulling data into the application. A simplified version of my current "pattern" looks like this (in psuedo-C# written in notepad): // This interface is implemented by all classes that needs the data interface IDataSubscriber { void RegisterData(Entity data); } // This interface is implemented by the data access class interface IDataProvider { void EnsureLoaded(Key dataKey); void RegisterSubscriber(IDataSubscriber subscriber); } class MyClassThatNeedsData : IDataSubscriber { IDataProvider _provider; MyClassThatNeedsData(IDataProvider provider) { _provider = provider; _provider.RegisterSubscriber(this); } public void RegisterData(Entity data) { // Save data for later StoreDataInCache(data); } void UseData(Key key) { // Make sure that the data has been stored in cache _provider.EnsureLoaded(key); Entity data = GetDataFromCache(key); } } class MyDataProvider : IDataProvider { List<IDataSubscriber> _subscribers; // Make sure that the data for key has been loaded to all subscribers public void EnsureLoaded(Key key) { if (HasKeyBeenMarkedAsLoaded(key)) return; PublishDataToSubscribers(key); MarkKeyAsLoaded(key); } // Force all subscribers to get a new version of the data for key public void ForceReload(Key key) { PublishDataToSubscribers(key); MarkKeyAsLoaded(key); } void PublishDataToSubscribers(Key key) { Entity data = FetchDataFromStore(key); foreach(var subscriber in _subscribers) { subscriber.RegisterData(data); } } } // This class will be spun off on startup and should make sure that all data is // preloaded as quickly as possible class MyPreloadingThread { IDataProvider _provider; MyPreloadingThread(IDataProvider provider) { _provider = provider; } void RunInBackground() { IEnumerable<Key> allKeys = GetAllKeys(); foreach(var key in allKeys) { _provider.EnsureLoaded(key); } } } I have a feeling though that this is not necessarily the best way of doing this.. Just the fact that explaining it seems to take two pages feels like an indication.. Any ideas? Any patterns out there I should have a look at?

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• Webdriver: Tests crash with internet explorer7 with error Modal dialog present

  - by user1207450

  Following tests is automated by using java and selenium-server-standalone-2.20.0.jar. The test crashes with the error: Page title is: cheese! - Google Search Starting browserTest 2922 [main] INFO org.apache.http.impl.client.DefaultHttpClient - I/O exception (org.apache.http.NoHttpResponseException) caught when processing request: The target server failed to respond 2922 [main] INFO org.apache.http.impl.client.DefaultHttpClient - Retrying request Exception in thread "main" org.openqa.selenium.UnhandledAlertException: Modal dialog present (WARNING: The server did not provide any stacktrace information) Command duration or timeout: 1.20 seconds Build info: version: '2.20.0', revision: '16008', time: '2012-02-27 19:03:04' System info: os.name: 'Windows XP', os.arch: 'x86', os.version: '5.1', java.version: '1.6.0_24' Driver info: driver.version: InternetExplorerDriver at sun.reflect.NativeConstructorAccessorImpl.newInstance0(Native Method) at sun.reflect.NativeConstructorAccessorImpl.newInstance(NativeConstructorAccessorImpl.java:39) at sun.reflect.DelegatingConstructorAccessorImpl.newInstance(DelegatingConstructorAccessorImpl.java:27) at java.lang.reflect.Constructor.newInstance(Constructor.java:513) at org.openqa.selenium.remote.ErrorHandler.createThrowable(ErrorHandler.java:170) at org.openqa.selenium.remote.ErrorHandler.throwIfResponseFailed(ErrorHandler.java:129) at org.openqa.selenium.remote.RemoteWebDriver.execute(RemoteWebDriver.java:438) at org.openqa.selenium.remote.RemoteWebDriver.startSession(RemoteWebDriver.java:139) at org.openqa.selenium.ie.InternetExplorerDriver.setup(InternetExplorerDriver.java:91) at org.openqa.selenium.ie.InternetExplorerDriver.<init>(InternetExplorerDriver.java:48) at com.pwc.test.java.InternetExplorer7.browserTest(InternetExplorer7.java:34) at com.pwc.test.java.InternetExplorer7.main(InternetExplorer7.java:27) Test Class: package com.pwc.test.java; import org.openqa.selenium.By; import org.openqa.selenium.WebDriver; import org.openqa.selenium.WebDriverBackedSelenium; import org.openqa.selenium.WebElement; import org.openqa.selenium.htmlunit.HtmlUnitDriver; import org.openqa.selenium.ie.InternetExplorerDriver; import com.thoughtworks.selenium.Selenium; public class InternetExplorer7 { /** * @param args */ public static void main(String[] args) { // TODO Auto-generated method stub WebDriver webDriver = new HtmlUnitDriver(); webDriver.get("http://www.google.com"); WebElement webElement = webDriver.findElement(By.name("q")); webElement.sendKeys("cheese!"); webElement.submit(); System.out.println("Page title is: "+webDriver.getTitle()); browserTest(); } public static void browserTest() { System.out.println("Starting browserTest"); String baseURL = "http://www.mail.yahoo.com"; WebDriver driver = new InternetExplorerDriver(); driver.get(baseURL); Selenium selenium = new WebDriverBackedSelenium(driver, baseURL); selenium.windowMaximize(); WebElement username = driver.findElement(By.id("username")); WebElement password = driver.findElement(By.id("passwd")); WebElement signInButton = driver.findElement(By.id(".save")); username.sendKeys("myusername"); password.sendKeys("magic"); signInButton.click(); driver.close(); } } I don't see any modal dialog when I launched the IE7/8 browser manually. What could be causing this?

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• paypal ipn working but stopping at 'thank you' page.

  - by Tarique Imam

  I am using the code for controller(CODEIGNITER): function paypal_tran(){ if (empty($_GET['action'])){ $_GET['action'] = 'process';} if($this-uri-segment ( 3 )){ $action=$this-uri-segment ( 3 ); } else{ $action='process'; } $ammount=39.99; $this-lenders_model-paypal_process($action,$this_script,$ammount); $this-load-view('view_paypal_tran'); } function ipn(){ if ($this->paypal_class->validate_ipn()) { $data = array( 'fname'=> 'fname', /* insert the user id */ 'lname'=>'lname' ); //$this->db->insert('ajax_test',$data); // For this example, we'll just email ourselves ALL the data. $subject = 'Instant Payment Notification - Recieved Payment'; $to = '[email protected]'; // your email $body = "An instant payment notification was successfully recieved\n"; $body .= "from ".$p->ipn_data['payer_email']." on ".date('m/d/Y'); $body .= " at ".date('g:i A')."\n\nDetails:\n"; foreach ($this->paypal_class->ipn_data as $key => $value) { $body .= "\n$key: $value"; } mail($to, $subject, $body); } } function success() { $this->load->view('paypal_succ_view'); } AND this is my model: function paypal_process($action,$this_script,$ammount){ switch ($action) { case 'process': // Process and order... // There should be no output at this point. To process the POST data, // the submit_paypal_post() function will output all the HTML tags which // contains a FORM which is submited instantaneously using the BODY onload // attribute. In other words, don't echo or printf anything when you're // going to be calling the submit_paypal_post() function. // This is where you would have your form validation and all that jazz. // You would take your POST vars and load them into the class like below, // only using the POST values instead of constant string expressions. // For example, after ensureing all the POST variables from your custom // order form are valid, you might have: // // $p->add_field('first_name', $_POST['first_name']); // $p->add_field('last_name', $_POST['last_name']); $this->paypal_class->add_field('business', '[email protected]'); $this->paypal_class->add_field('return', $this_script.'/success'); $this->paypal_class->add_field('cancel_return', $this_script.'/cancel'); $this->paypal_class->add_field('notify_url', $this_script.'/ipn'); $this->paypal_class->add_field('item_name', 'Lenders Account for one month'); $this->paypal_class->add_field('amount', $ammount); $this->paypal_class->submit_paypal_post(); // submit the fields to paypal $this->paypal_class->dump_fields(); // for debugging, output a table of all the fields break; PROBLEM IS IPN IS NOT WORKING. THE HIDDEN FIELD HAS VALUE FOR REDIRECT TO IPN, BUT NOT WORKING!!PLS HELP

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• Algorithm to Render a Horizontal Binary-ish Tree in Text/ASCII form

  - by Justin L.

  It's a pretty normal binary tree, except for the fact that one of the nodes may be empty. I'd like to find a way to output it in a horizontal way (that is, the root node is on the left and expands to the right). I've had some experience expanding trees vertically (root node at the top, expanding downwards), but I'm not sure where to start, in this case. Preferably, it would follow these couple of rules: If a node has only one child, it can be skipped as redundant (an "end node", with no children, is always displayed) All nodes of the same depth must be aligned vertically; all nodes must be to the right of all less-deep nodes and to the left of all deeper nodes. Nodes have a string representation which includes their depth. Each "end node" has its own unique line; that is, the number of lines is the number of end nodes in the tree, and when an end node is on a line, there may be nothing else on that line after that end node. As a consequence of the last rule, the root node should be in either the top left or the bottom left corner; top left is preferred. For example, this is a valid tree, with six end nodes (node is represented by a name, and its depth): [a0]------------[b3]------[c5]------[d8] \ \ \----------[e9] \ \----[f5] \--[g1]--------[h4]------[i6] \ \--------------------[j10] \-[k3] Which represents the horizontal, explicit binary tree: 0 a / \ 1 g * / \ \ 2 * * * / \ \ 3 k * b / / \ 4 h * * / \ \ \ 5 * * f c / \ / \ 6 * i * * / / \ 7 * * * / / \ 8 * * d / / 9 * e / 10 j (branches folded for compactness; * representing redundant, one-child nodes; note that *'s are actual nodes, storing one child each, just with names omitted here for presentation sake) (also, to clarify, I'd like to generate the first, horizontal tree; not this vertical tree) I say language-agnostic because I'm just looking for an algorithm; I say ruby because I'm eventually going to have to implement it in ruby anyway. Assume that each Node data structure stores only its id, a left node, and a right node. A master Tree class keeps tracks of all nodes and has adequate algorithms to find: A node's nth ancestor A node's nth descendant The generation of a node The lowest common ancestor of two given nodes Anyone have any ideas of where I could start? Should I go for the recursive approach? Iterative?

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• Undefined Behavior and Sequence Points Reloaded

  - by Nawaz

  Consider this topic a sequel of the following topic: Previous Installment Undefined Behavior and Sequence Points Let's revisit this funny and convoluted expression (the italicized phrases are taken from the above topic *smile* ): i += ++i; We say this invokes undefined-behavior. I presume that when say this, we implicitly assume that type of i is one of built-in types. So my question is: what if the type of i is a user-defined type? Say it's type is Index which is defined later in this post (see below). Would it still invoke undefined-behavior? If yes, why? Is it not equivalent to writing i.operator+=(i.operator ++()); or even syntactically simpler i.add(i.inc());? Or, do they too invoke undefined-behavior? If no, why not? After all, the object i gets modified twice between consecutive sequence points. Please recall the rule of thumb : an expression can modify an object's value only once between consecutive "sequence points. And if i += ++i is an expression, then it must invoke undefined-behavior. If so, then it's equivalents i.operator+=(i.operator ++()); and i.add(i.inc()); must also invoke undefined-behavior which seems to be untrue! (as far as I understand) Or, i += ++i is not an expression to begin with? If so, then what is it and what is the definition of expression? If it's an expression, and at the same time, it's behavior is also well-defined, then it implies that number of sequence points associated with an expression somehow depends on the type of operands involved in the expression. Am I correct (even partly)? By the way, how about this expression? a[++i] = i; //taken from the previous topic. but here type of `i` is Index. class Index { int state; public: Index(int s) : state(s) {} Index& operator++() { state++; return *this; } Index& operator+=(const Index & index) { state+= index.state; return *this; } operator int() { return state; } Index & add(const Index & index) { state += index.state; return *this; } Index & inc() { state++; return *this; } };

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• Classifying captured data in unknown format?

  - by monch1962

  I've got a large set of captured data (potentially hundreds of thousands of records), and I need to be able to break it down so I can both classify it and also produce "typical" data myself. Let me explain further... If I have the following strings of data: 132T339G1P112S 164T897F5A498S 144T989B9B223T 155T928X9Z554T ... you might start to infer the following: possibly all strings are 14 characters long the 4th, 8th, 10th and 14th characters may always be alphas, while the rest are numeric the first character may always be a '1' the 4th character may always be the letter 'T' the 14th character may be limited to only being 'S' or 'T' and so on... As you get more and more samples of real data, some of these "rules" might disappear; if you see a 15 character long string, then you have evidence that the 1st "rule" is incorrect. However, given a sufficiently large sample of strings that are exactly 14 characters long, you can start to assume that "all strings are 14 characters long" and assign a numeric figure to your degree of confidence (with an appropriate set of assumptions around the fact that you're seeing a suitably random set of all possible captured data). As you can probably tell, a human can do a lot of this classification by eye, but I'm not aware of libraries or algorithms that would allow a computer to do it. Given a set of captured data (significantly more complex than the above...), are there libraries that I can apply in my code to do this sort of classification for me, that will identify "rules" with a given degree of confidence? As a next step, I need to be able to take those rules, and use them to create my own data that conforms to these rules. I assume this is a significantly easier step than the classification, but I've never had to perform a task like this before so I'm really not sure how complex it is. At a guess, Python or Java (or possibly Perl or R) are possibly the "common" languages most likely to have these sorts of libraries, and maybe some of the bioinformatic libraries do this sort of thing. I really don't care which language I have to use; I need to solve the problem in whatever way I can. Any sort of pointer to information would be very useful. As you can probably tell, I'm struggling to describe this problem clearly, and there may be a set of appropriate keywords I can plug into Google that will point me towards the solution. Thanks in advance

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• CSS precedence order? My lecture slides are correct or not?

  - by Michael Mao

  Hi all: I've noticed that there are a couple of similar questions and answers at SO already, but let me clarify my specific question here first: I've got lecture slides which states like this: To be frank, I haven't heard of this rule of css precedence myself, and I googled to find something with similar topic but not quite like that : here To have a test myself, I've made a test page on my own server here After running it on FireFox 3.6.3, I am sure it does not show the way as it should be, according to the statement in lecture slides: imported stylesheet ? am I doing it wrong? I cannot see its effect using FireBug it says that embedded stylesheet has a higher precedence over linked/imported stylesheets, however, it doesn't work, if I put the linked/imported tag AFTER that. inline style vs html attributes ? I've got an image where I firstly set its inline style to control the width and height, then use direct html attributes width/height to try modifying that, but failed... Below is the source code : <html> <head> <style type="text/css"> #target { border : 2px solid green; color : green; } </style> <link rel="stylesheet" href="./linked.css" type="text/css" media="screen" /> </head> <body> <div id="target">A targeted div tag on page.</div> <img src="cat.jpg" alt="" style="width : 102px; height : 110px;" width="204px" height="220px" /> </body> </html> Can any experienced CSS guys help me figure out if the slide is correct or not? Frankly speaking, I am puzzled myself, as I can clearly see some other "incorrect" statements here and there amongst the slides, such as JavaScript is on client-side (how about server-side JavaScript?) and "Embedded styles are in the head section of a web page "(what the heck? I am not allowed to put it inside the body tag?) Sorry about this silly question, the exam is on TOMORROW, and I now see a lot of things to think about :)

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• difference equations in MATLAB - why the need to switch signs?

  - by jefflovejapan

  Perhaps this is more of a math question than a MATLAB one, not really sure. I'm using MATLAB to compute an economic model - the New Hybrid ISLM model - and there's a confusing step where the author switches the sign of the solution. First, the author declares symbolic variables and sets up a system of difference equations. Note that the suffixes "a" and "2t" both mean "time t+1", "2a" means "time t+2" and "t" means "time t": %% --------------------------[2] MODEL proc-----------------------------%% % Define endogenous vars ('a' denotes t+1 values) syms y2a pi2a ya pia va y2t pi2t yt pit vt ; % Monetary policy rule ia = q1*ya+q2*pia; % ia = q1*(ya-yt)+q2*pia; %%option speed limit policy % Model equations IS = rho*y2a+(1-rho)yt-sigma(ia-pi2a)-ya; AS = beta*pi2a+(1-beta)*pit+alpha*ya-pia+va; dum1 = ya-y2t; dum2 = pia-pi2t; MPs = phi*vt-va; optcon = [IS ; AS ; dum1 ; dum2; MPs]; He then computes the matrix A: %% ------------------ [3] Linearization proc ------------------------%% % Differentiation xx = [y2a pi2a ya pia va y2t pi2t yt pit vt] ; % define vars jopt = jacobian(optcon,xx); % Define Linear Coefficients coef = eval(jopt); B = [ -coef(:,1:5) ] ; C = [ coef(:,6:10) ] ; % B[c(t+1) l(t+1) k(t+1) z(t+1)] = C[c(t) l(t) k(t) z(t)] A = inv(C)*B ; %(Linearized reduced form ) As far as I understand, this A is the solution to the system. It's the matrix that turns time t+1 and t+2 variables into t and t+1 variables (it's a forward-looking model). My question is essentially why is it necessary to reverse the signs of all the partial derivatives in B in order to get this solution? I'm talking about this step: B = [ -coef(:,1:5) ] ; Reversing the sign here obviously reverses the sign of every component of A, but I don't have a clear understanding of why it's necessary. My apologies if the question is unclear or if this isn't the best place to ask.

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• WPF Application Slow Unresponsive when demonstrating using remote sharing software

  - by Kev

  After spending 14 hours on this I think its time to share my woes and see if anyone has experienced this issue before. Ill describe the issue and tests I have done to rule out certain things. Ok so I have a WPF application which loads in data from an SQL database. I am using DevExpress Components for datagrids, ribbons etc.. and FluentNhibernate to provide a session for database operations. I am also using log4net to log events to a textfile. Using the application on my laptop with SQL Express 2008 works fine.. the application starts up, retrieves 1000 records and I can tab through the controls on the ribbon. Now, I decided to demo the application to a third party and used remote login/sharing software online to share my desktop with the other person so as I could load the application on my laptop and they could view me using the application. Now, the application takes approx 45 seconds to load... 30 seconds with a blank database where as, when im not sharing out my screen using the online software the application loads in about 7-10 seconds. As well as that, even using the controls in the application during the demo were very sticky, slow and unresponsive. During the sharing session though however I was able to use other applications without any problems.. everything else worked fine. But I cannot understand how my application works ok under normal conditions , even browsing the net at the same time etc... BUT totally fails to perform correctly when I am sharing a session with another user... the CPU usage shot up to 100% too at times when the application was trying to start up... Please see below a list of 3rd party dlls I am using as references in my project. DevExpress dlls FluidKit PixelLab.WPF PixelLab.Common Galasoft WPF Kit FluentNHibernate NHibernate Nhibernate.ByteCode.Castle Skype4ComLib TXTEXTControl log4net LinqKit All of these DLLs are in the output folder with the application dlls created from the class assemblys in the project. So when installed via an installer on a machine the dlls will be in the same application folder as the application file itself. Many thanks

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• Calculating a Sample Covariance Matrix for Groups with plyr

  - by John A. Ramey

  I'm going to use the sample code from http://gettinggeneticsdone.blogspot.com/2009/11/split-apply-and-combine-in-r-using-plyr.html for this example. So, first, let's copy their example data: mydata=data.frame(X1=rnorm(30), X2=rnorm(30,5,2), SNP1=c(rep("AA",10), rep("Aa",10), rep("aa",10)), SNP2=c(rep("BB",10), rep("Bb",10), rep("bb",10))) I am going to ignore SNP2 in this example and just pretend the values in SNP1 denote group membership. So then, I may want some summary statistics about each group in SNP1: "AA", "Aa", "aa". Then if I want to calculate the means for each variable, it makes sense (modifying their code slightly) to use: > ddply(mydata, c("SNP1"), function(df) data.frame(meanX1=mean(df$X1), meanX2=mean(df$X2))) SNP1 meanX1 meanX2 1 aa 0.05178028 4.812302 2 Aa 0.30586206 4.820739 3 AA -0.26862500 4.856006 But what if I want the sample covariance matrix for each group? Ideally, I would like a 3D array, where the I have the covariance matrix for each group, and the third dimension denotes the corresponding group. I tried a modified version of the previous code and got the following results that have convinced me that I'm doing something wrong. > daply(mydata, c("SNP1"), function(df) cov(cbind(df$X1, df$X2))) , , = 1 SNP1 1 2 aa 1.4961210 -0.9496134 Aa 0.8833190 -0.1640711 AA 0.9942357 -0.9955837 , , = 2 SNP1 1 2 aa -0.9496134 2.881515 Aa -0.1640711 2.466105 AA -0.9955837 4.938320 I was thinking that the dim() of the 3rd dimension would be 3, but instead, it is 2. Really this is a sliced up version of the covariance matrix for each group. If we manually compute the sample covariance matrix for aa, we get: [,1] [,2] [1,] 1.4961210 -0.9496134 [2,] -0.9496134 2.8815146 Using plyr, the following gives me what I want in list() form: > dlply(mydata, c("SNP1"), function(df) cov(cbind(df$X1, df$X2))) $aa [,1] [,2] [1,] 1.4961210 -0.9496134 [2,] -0.9496134 2.8815146 $Aa [,1] [,2] [1,] 0.8833190 -0.1640711 [2,] -0.1640711 2.4661046 $AA [,1] [,2] [1,] 0.9942357 -0.9955837 [2,] -0.9955837 4.9383196 attr(,"split_type") [1] "data.frame" attr(,"split_labels") SNP1 1 aa 2 Aa 3 AA But like I said earlier, I would really like this in a 3D array. Any thoughts on where I went wrong with daply() or suggestions? Of course, I could typecast the list from dlply() to a 3D array, but I'd rather not do this because I will be repeating this process many times in a simulation. As a side note, I found one method (http://www.mail-archive.com/[email protected]/msg86328.html) that provides the sample covariance matrix for each group, but the outputted object is bloated. Thanks in advance.

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• Error trying to run rails server

  - by David87

  I am trying to get a basic Rails application to run on my Mac OS X 10.6.5. I created a new app called demo (rails new demo), then went into the demo directory and tried to start the app with rails server. Here is the error message I received: "/Users/dpetrovi/.gem/ruby/1.8/gems/sqlite3-ruby-1.3.2/lib/sqlite3/sqlite3_native.bundle: [BUG] Segmentation fault ruby 1.8.7 (2010-12-23 patchlevel 330) [i686-darwin10] Abort trap" I checked bundle install in the demo folder: "Using rake (0.8.7) Using abstract (1.0.0) Using activesupport (3.0.3) Using builder (2.1.2) Using i18n (0.5.0) Using activemodel (3.0.3) Using erubis (2.6.6) Using rack (1.2.1) Using rack-mount (0.6.13) Using rack-test (0.5.6) Using tzinfo (0.3.23) Using actionpack (3.0.3) Using mime-types (1.16) Using polyglot (0.3.1) Using treetop (1.4.9) Using mail (2.2.13) Using actionmailer (3.0.3) Using arel (2.0.6) Using activerecord (3.0.3) Using activeresource (3.0.3) Using bundler (1.0.7) Using thor (0.14.6) Using railties (3.0.3) Using rails (3.0.3) Using sqlite3-ruby (1.3.2) Your bundle is complete! Use bundle show [gemname] to see where a bundled gem is installed." Ruby, RubyGems, and sqlite3 were installed using MacPorts. Then I used gem to try to install the sqlite3-ruby interface. (sudo gem install sqlite3-ruby). Here is where I first noticed something could be off: "Successfully installed sqlite3-ruby-1.3.2 1 gem installed Installing ri documentation for sqlite3-ruby-1.3.2... No definition for libversion Enclosing class/module 'mSqlite3' for class Statement not known Installing RDoc documentation for sqlite3-ruby-1.3.2... No definition for libversion Enclosing class/module 'mSqlite3' for class Statement not known " I had rails running well on my system a few months ago, so I figured maybe I had some duplicates and it was trying to use the wrong one. I ran: "for cmd in ruby irb gem rake; do which $cmd; done" and got: "/opt/local/bin/ruby /opt/local/bin/irb /opt/local/bin/gem /opt/local/bin/rake" Checking where sqlite3 also gets me: "/opt/local/bin/sqlite3" so they all seem to be in the right place. Obviously /opt/local/bin is in my system path. If I check gems server, it shows that I have installed sqlite3-ruby 1.3.2 gem. Not sure what the problem could be? I am using ruby 1.8.7 (2010-12-23 patchlevel 330) [i686-darwin10]. Macports claims this is the latest (although ive seen 1.9.1) One more thing-- in irb, I tried to check which version of sqlite3 my sqlite3-ruby is bound to, but I can only get this far: ":irb(main):001:0 require 'rubygems' = true irb(main):002:0 require 'sqlite3' /Users/dpetrovi/.gem/ruby/1.8/gems/sqlite3-ruby-1.3.2/lib/sqlite3/sqlite3_native.bundle: [BUG] Segmentation fault ruby 1.8.7 (2010-12-23 patchlevel 330) [i686-darwin10] Abort trap" Any suggestions? Im hoping I overlooked something obvious. Thanks

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• jquery two ajax call asynchrounsly in asp.net not working...

  - by eswaran

  Hi all, I am developed an web application in asp.net. In this application I have used jquery ajax for some pages. In this application, when I make two ajax call asynchrounoulsy that would not do as I expceted. what is happening is even the second ajax call finishes i can see the result when the maximum time out ajax call finished. I mean i can see the both results in the same time, not one by one. for an example. I have 3 pages 1) main.aspx - for make two ajax request. 2) totalCount.aspx - to find the total count. (max it takes 7 seconds to return, as corresponding table contains 3 lak records) 3) rowCount.aspx - to find the row details. (max it takes 5 seconds to return result). due to this scene, I have planed to make asyn call in jquery ajax in asp.net. here is my code... function getResult() { getTotalCount(); getRows(); } // it takes max 7 seconds to complete // as it take 7 seconds it should display second.( I mean after the rows dispaying) // but displaying both at the same time after the max time consuming ajax call completed. function getTotalCount() { $.ajax({ type : "POST", async : true, url : "totalCount.aspx?data1=" + document.getElementById("data").value, success : function(responseText) { $("#totalCount").attr("value", responseText); } }) } // it takes max 5 seconds to complete. // after finished, this should display first.( i mean before total count displays) // but displaying both at the same time after the max time consuming ajax call completed. function getRows() { $.ajax({ type : "POST", url : "getrows.aspx?data1=" + document.getElementById("data").value, async : true, success : function(responseText) { $("#getRows").attr("value", responseText); } }); } I would like to know, If there is any possible to make asyn call in jquery ajax in asp.net. I searched in net, I got some points that says we cannot do this in asp.net ref link: http://www.mail-archive.com/[email protected]/msg55125.html if we can do this in asp.net How to do that? thanks r.eswaran.

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• Checking if an int is prime more efficiently

  - by SipSop

  I recently was part of a small java programming competition at my school. My partner and I have just finished our first pure oop class and most of the questions were out of our league so we settled on this one (and I am paraphrasing somewhat): "given an input integer n return the next int that is prime and its reverse is also prime for example if n = 18 your program should print 31" because 31 and 13 are both prime. Your .class file would then have a test case of all the possible numbers from 1-2,000,000,000 passed to it and it had to return the correct answer within 10 seconds to be considered valid. We found a solution but with larger test cases it would take longer than 10 seconds. I am fairly certain there is a way to move the range of looping from n,..2,000,000,000 down as the likely hood of needing to loop that far when n is a low number is small, but either way we broke the loop when a number is prime under both conditions is found. At first we were looping from 2,..n no matter how large it was then i remembered the rule about only looping to the square root of n. Any suggestions on how to make my program more efficient? I have had no classes dealing with complexity analysis of algorithms. Here is our attempt. public class P3 { public static void main(String[] args){ long loop = 2000000000; long n = Integer.parseInt(args[0]); for(long i = n; i<loop; i++) { String s = i +""; String r = ""; for(int j = s.length()-1; j>=0; j--) r = r + s.charAt(j); if(prime(i) && prime(Long.parseLong(r))) { System.out.println(i); break; } } System.out.println("#"); } public static boolean prime(long p){ for(int i = 2; i<(int)Math.sqrt(p); i++) { if(p%i==0) return false; } return true; } } ps sorry if i did the formatting for code wrong this is my first time posting here. Also the output had to have a '#' after each line thats what the line after the loop is about Thanks for any help you guys offer!!!

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• Finding underlying cause of Window 7 Account corruption.

  - by Carl Jokl

  I have been having trouble with my Sister's computer which I built. It is running Windows 7 Ultimate x64. The problem is that I have had problems with the accounts becoming corrupted. First problems manifest themselves in the form of Windows saying the profile failed to be loaded properly and a temporary profile. Eventually the account will not allow login at all. An error message along the lines the authentication service failing the login. I have found information about this problem and how to fix it. The problem being that something has corrupted the account profile and backing up and recreating the accounts fixes the problem. I have been able to fix things and get logins working again but over the period of usually about a week it happens again. Bit by bit the accounts corrupt and then it is back to square one. I am frustrated because I don't know what the underlying cause of the problem is i.e. what is causing the accounts to be corrupted in the first place. At the moment I am just treating the symptoms. I was hoping someone who may have more experience with dealing with this problem might be able to help me find the root cause. Some articles suggest that Norton Internet Security is a big culprit of this problem which is installed. I could try uninstalling Norton and see if it helps. The one thing which is different about this computer to any other I have built is that it has a solid state drive. Actually it has both a hard drive and solid state drive. The documents and settings i.e. the Users directory is stored on the hard drive. This was done following an article about moving the user account data onto a separate drive on Windows 7 which I found on the Internet. Moving the User accounts is more of a pain under Windows 7 and this solution involved creating a low level file system link to the folder from the boot drive (Solid State) to the Hard Drive. The idea is that the computer behaves just as if it is accessing the User's folder from the boot drive but actually the data is stored on the hard drive. This may have nothing to do with the cause of the problem but due to the problem being user account corruption it is a possibility I have not been able to rule out. Any help would be appreciated as I would be glad to see the back of this problem.

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• Testing for existence using SELECT WHERE HAVING and NOT HAVING in a grouped subset

  - by IanC

  I have data on which I need to count +1 if a particular condition exists or another condition doesn't exist. I'm using SQL Server 2008. I shred the following simplified sample XML into a temp table and validate it: <product type="1"> <param type="1"> <item mode="0" weight="1" /> </param> <param type="2"> <item mode="1" weight="1" /> <item mode="0" weight="0.1" /> </param> <param type="3"> <item mode="1" weight="0.75" /> <item mode="1" weight="0.25" /> </param> </product> The validation in concern is the following rule: For each product type, for each param type, mode may be 0 & (1 || 2). In other words, there may be a 0(s), but then 1s or 2s are required, or there may be only 1(s) or 2(s). There cannot be only 0s, and there cannot be 1s and 2s. The only part I haven't figured out is how to detect if there are only 0s. This seems like a "not having" problem. The validation code (for this part): WITH t1 AS ( SELECT SUM(t.ParamWeight) AS S, COUNT(1) AS C, t.ProductTypeID, t.ParamTypeID, t.Mode FROM @t AS t GROUP BY t.ProductTypeID, t.ParamTypeID, t.Mode ), ... UNION ALL SELECT TOP (1) 1 -- only mode 0 & (1 || 2) is allowed FROM t1 WHERE t1.Mode IN (1, 2) GROUP BY t1.ProductTypeID, t1.ParamTypeID HAVING COUNT(1) > 1 UNION ALL ... ) SELECT @C = COUNT(1) FROM t2 This will show if any mode 1s & 2s are mixed, but not if the group contains only a 0. I'm sure there is a simple solution, but it's evading me right now. EDIT: I thought of a "cheat" that works perfectly. I added the following to the above: SELECT TOP (1) 1 -- only mode 0 & (null || 1 || 2) is allowed FROM t1 GROUP BY t1.ProductTypeID, t1.ParamTypeID HAVING SUM(t1.Mode) = 0 However, I'd still like to know how to do this without cheating.

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• C# Detect Localhost Port Usage

  - by ThaKidd

  In advance, thank you for your advice. I am currently working on a program which uses Putty to create a SSH connection with a server that uses local port forwarding to enable a client, running my software, to access the service behind the SSH server via localhost. IE: client:20100 - Internet - Remote SSH server exposed via router/firewall - Local Intranet - Intranet Web POP3 Server:110. Cmd Line: "putty -ssh -2 -P 22 -C -L 20100:intranteIP:110 -pw sshpassword sshusername@sshserver" Client would use putty to create a SSH connection with the SSH server specifying in the connection string that it would like to tie port 110 of the Intranet POP3 Server to port 20100 on the client system. Therefore the client would be able to open up a mail client to localhost:20100 and interact with the Internal POP3 server over the SSH tunnel. The above is a general description. I already know what I am trying to do will work without a problem so am not looking for debate on the above. The question is this...How can I ensure the local port (I cannot use dynamic ports, so it must be static) on localhost is not being used or listened to by any other application? I am currently executing this code in my C# app: private bool checkPort(int port) { try { //Create a socket on the current IPv4 address Socket TestSocket = new Socket(AddressFamily.InterNetwork, SocketType.Stream, ProtocolType.Tcp); // Create an IP end point IPEndPoint localIP = new IPEndPoint(IPAddress.Parse("127.0.0.1"), port); // Bind that port TestSocket.Bind(localIP); // Cleanup TestSocket.Close(); return false; } catch (Exception e) { // Exception occurred. Port is already bound. return true; } } I am currently calling this function starting with a specific port in a for loop to get the 'false' return at the first available port. The first port I try is actually being listened to by uTorrent. The above code does not catch this and my connection fails. What is the best method to ensure a port is truly free? I do understand some other program may grab the port during/after I have tested it. I just need to find something that will ensure it is not currently in use AT ALL when the test is executed. If there is a way to truly reserve the localhost port during the test, I would love to hear about it.

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• PHP email form multiple select

  - by Justin Goodman

  I'm trying to set up a simple PHP contact form for a website and I need some help modifying the PHP to list multiple items from a select menu and would appreciate the help. I'm a graphic designer, not a developer, so a lot of this is way over my head. This is the problem area here: <label for="Events[]">Which Event(s) Will You Be Attending?</label> <div class="input-bg"> <select name="Events[]" size="6" multiple="MULTIPLE" class="required" id="Events[]"> <option value="Wednesday">Portfolio Show June 16</option> <option value="Thursday">Portfolio Show June 17</option> <option value="Saturday">Graduation Ceremony</option> <option value="Saturday Eve">Graduation Party</option> <option value="Not Sure">Not Sure</option> <option value="Not Coming">Not Coming</option> </select> </div> And here's the PHP: <?php // CHANGE THE VARIABLES BELOW $EmailFrom = "[email protected]"; $EmailTo = "[email protected]"; $Subject = "Graduation RSVP"; $Name = Trim(stripslashes($_POST['Name'])); $Email = Trim(stripslashes($_POST['Email'])); $Guests = Trim(stripslashes($_POST['Guests'])); $Events = Trim(stripslashes($_POST['Events'])); // prepare email body text $Body = ""; $Body .= "Name: "; $Body .= $Name; $Body .= "\n"; $Body .= "Email: "; $Body .= $Email; $Body .= "\n"; $Body .= "Guests: "; $Body .= $Guests; $Body .= "\n"; $Body .= "Events: "; $Body .= $Events; $Body .= "\n"; // send email $success = mail($EmailTo, $Subject, $Body, "From: <$EmailFrom>"); // redirect to success page // CHANGE THE URL BELOW TO YOUR "THANK YOU" PAGE if ($success){ print "<meta http-equiv=\"refresh\" content=\"0;URL=http://justgooddesign.net/graduation\">"; } else{ print "<meta http-equiv=\"refresh\" content=\"0;URL=http://justgooddesign.net/graduation/error.html\">"; } ?> Any help really is appreciated!

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