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• Adding a mini admin to a webpage.

  - by DADU

  Hello Picture this: you are creating a little module that people can incorporate into their website easily, for example, a little contact form. It would consist of a PHP file that outputs some HTML, a Javascript file (ajax etc.), a CSS file and a CSS skin. Now the person who doesn't know much about coding wants to integrate it on a webpage (website/index.php). We could do this with three rules of code: <link rel="stylesheet" href="module/css/module.css" /> <script src="module/js/module.js"></script> <?php require_once 'module/module.php'; ?> There's no doubt this part is questionable, right? Now when we want to add an admin for this little module, there are two options: Accessing the admin via an extra URL like website/module/admin.php and after authentication, displaying a page where the person can do all the settings. The person then goes back to index.php to see the results. Enabling the admin via an extra URL like website/module/admin.php and after authentication, redirecting back to index.php. The person can now edit the module directly (HTML5 contenteditable) and see changes live, on the webpage where everybody else will see it when the person saves the changes. Option 2 has a couple of advantages: The person doesn't have to toggle between admin and index.php. The person can see directly how it's looking at the webpage it's integrated in. The person probably feels like the module is more part of the webpage/website. Of course option 2 has some disadvantages too: Not everything works well editing it inline. The person would need to have an HTML5 compliant browser. Probably some more I can't think of right now. Now I have a few concerns that's I can't seem to see a clear answer to. How would we let the person integrate the admin on their webpage? The admin files only need to be included in index.php if the person has choosen to edit the module via the url (website/module/admin.php). But how can we do this if we have a admin.css file that belongs in the head section, an admin.php file that goes into the body, and another admin.js file that's included at the end of the body? How would we know the file that admin.php needs to redirect back to, after authentication? index.php could be any webpage with any name. Any real life website/web apps examples using this principle are welcome too. If there's something unclear, I am glad to add additional info.

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• How do I get the next token in a Cstring if I want to use it as an int? (c++)

  - by Van

  My objective is to take directions from a user and eventually a text file to move a robot. The catch is that I must use Cstrings(such as char word[];) rather than the std::string and tokenize them for use. the code looks like this: void Navigator::manualDrive() { char uinput[1]; char delim[] = " "; char *token; cout << "Enter your directions below: \n"; cin.ignore(); cin.getline (uinput, 256); token=strtok(uinput, delim); if(token == "forward") { int inches; inches=token+1; travel(inches); } } I've never used Cstrings I've never tokenized anything before, and I don't know how to write this. Our T.A.'s expect us to google and find all the answers because they are aware we've never been taught these methods. Everyone in my lab is having much more trouble than usual. I don't know the code to write but I know what I want my program to do. I want it to execute like this: 1) Ask for directions. 2) cin.getline the users input 3) tokenize the inputed string 4) if the first word token == "forward" move to the next token and find out how many inches to move forward then move forward 5) else if the first token == "turn" move to the next token. if the next token == "left" move to the next token and find out how many degrees to turn left I will have to do this for forward x, backward x, turn left x, turn right x, and stop(where x is in inches or degrees). I already wrote functions that tell the robot how to move forward an inch and turn in degrees. I just need to know how to convert the inputted strings to all lowercase letters and move from token to token and convert or extract the numbers from the string to use them as integers. If all is not clear you can read my lab write up at this link: http://www.cs.utk.edu/~cs102/robot_labs/Lab9.html If anything is unclear please let me know, and I will clarify as best I can.

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• Abstract mapping with custom JiBX marshaller

  - by aweigold

  I have created a custom JiBX marshaller and verified it works. It works by doing something like the following: <binding xmlns:tns="http://foobar.com/foo" direction="output"> <namespace uri="http://foobar.com/foo" default="elements"/> <mapping class="java.util.HashMap" marshaller="com.foobar.Marshaller1"/> <mapping name="context" class="com.foobar.Context"> <structure field="configuration"/> </mapping> </binding> However I need to create multiple marshallers for different HashMaps. So I tried to reference it with abstract mapping like this: <binding xmlns:tns="http://foobar.com/foo" direction="output"> <namespace uri="http://foobar.com/foo" default="elements"/> <mapping abstract="true" type-name="configuration" class="java.util.HashMap" marshaller="com.foobar.Marshaller1"/> <mapping abstract="true" type-name="overrides" class="java.util.HashMap" marshaller="com.foobar.Marshaller2"/> <mapping name="context" class="com.foobar.Context"> <structure map-as="configuration" field="configuration"/> <structure map-as="overrides" field="overrides"/> </mapping> </binding> However when doing so, when I attempt to build the binding, I receive the following: Error during code generation for file "E:\project\src\main\jibx\foo.jibx" - this may be due to an error in your binding or classpath, or to an error in the JiBX code My guess is that either I'm missing something I need to implement to enable my custom marshaller for abstract mapping, or custom marshallers do not support abstract mapping. I have found the IAbstractMarshaller interface in the JiBX API (http://jibx.sourceforge.net/api/org/jibx/runtime/IAbstractMarshaller.html), however the documentation seems unclear to me on if this is what I need to implement, as well as how it works if so. I have not been able to find an implementation of this interface to work off of as an example. My question is, how do you do abstract mapping with custom marshallers (if it's possible)? If it is done via the IAbstractMarshaller interface, how does it work and/or how should I implement it?

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• No matter what, I can't get this stupid progress bar to update from a thread!

  - by Synthetix

  I have a Windows app written in C (using gcc/MinGW) that works pretty well except for a few UI problems. One, I simply cannot get the progress bar to update from a thread. In fact, I probably can't get ANY UI stuff to update. Basically, I have a spawned thread that does some processing, and from that thread I attempt to update the progress bar in the main thread. I tried this by using PostMessage() to the main hwnd, but no luck even though I can do other things like open message boxes. However, it's unclear whether the message box is getting called within the thread or on the main thread. Here's some code: //in header/globally accessible HWND wnd; //main application window HWND progress_bar; //progress bar typedef struct { //to pass to thread DWORD mainThreadId; HWND mainHwnd; char *filename; } THREADSTUFF; //callback function LRESULT CALLBACK WndProc(HWND hwnd, UINT msg, WPARAM wParam, LPARAM lParam){ switch(msg){ case WM_CREATE:{ //create progress bar progress_bar = CreateWindowEx( 0, PROGRESS_CLASS, (LPCTSTR)NULL, WS_CHILD | WS_VISIBLE, 79,164,455,15, hwnd, (HMENU)20, NULL, NULL); break; } case WM_COMMAND:{ if(LOWORD(wParam)==2){ //do some processing in a thread //struct of stuff I need to pass to thread THREADSTUFF *threadStuff; threadStuff = (THREADSTUFF*)malloc(sizeof(*threadStuff)); threadStuff->mainThreadId = GetCurrentThreadId(); threadStuff->mainHwnd = hwnd; threadStuff->filename = (void*)&filename; hThread1 = CreateThread(NULL,0,convertFile (LPVOID)threadStuff,0,NULL); }else if(LOWORD(wParam)==5){ //update progress bar MessageBox(hwnd,"I got a message!", "Message", MB_OK | MB_ICONINFORMATION); PostMessage(progress_bar,PBM_STEPIT,0,CLR_DEFAULT); } break; } } } This all seems to work okay. The problem is in the thread: DWORD WINAPI convertFile(LPVOID params){ //get passed params, this works perfectly fine THREADSTUFF *tData = (THREADSTUFF*)params; MessageBox(tData->mainHwnd,tData->filename,"File name",MB_OK | MB_ICONINFORMATION); //yep PostThreadMessage(tData->mainThreadId,WM_COMMAND,5,0); //only shows message PostMessage(tData->mainHwnd,WM_COMMAND,5,0); //only shows message } When I say, "only shows message," that means the MessageBox() function in the callback works, but not the PostMessage() to update the position of the progress bar. What am I missing?

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• How to use unset() for this Linear Linked List in PHP

  - by Peter

  I'm writing a simple linear linked list implementation in PHP. This is basically just for practice... part of a Project Euler problem. I'm not sure if I should be using unset() to help in garbage collection in order to avoid memory leaks. Should I include an unset() for head and temp in the destructor of LLL? I understand that I'll use unset() to delete nodes when I want, but is unset() necessary for general clean up at any point? Is the memory map freed once the script terminates even if you don't use unset()? I saw this SO question, but I'm still a little unclear. Is the answer that you simply don't have to use unset() to avoid any sort of memory leaks associated with creating references? I'm using PHP 5.. btw. Unsetting references in PHP PHP references tutorial Here is the code - I'm creating references when I create $temp and $this-head at certain points in the LLL class: class Node { public $data; public $next; } class LLL { // The first node private $head; public function __construct() { $this->head = NULL; } public function insertFirst($data) { if (!$this->head) { // Create the head $this->head = new Node; $temp =& $this->head; $temp->data = $data; $temp->next = NULL; } else { // Add a node, and make it the new head. $temp = new Node; $temp->next = $this->head; $temp->data = $data; $this->head =& $temp; } } public function showAll() { echo "The linear linked list:<br/>&nbsp;&nbsp;"; if ($this->head) { $temp =& $this->head; do { echo $temp->data . " "; } while ($temp =& $temp->next); } else { echo "is empty."; } echo "<br/>"; } } Thanks!

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• Can you get a Func<T> (or similar) from a MethodInfo object?

  - by Dan Tao

  I realize that, generally speaking, there are performance implications of using reflection. (I myself am not a fan of reflection at all, actually; this is a purely academic question.) Suppose there exists some class that looks like this: public class MyClass { public string GetName() { return "My Name"; } } Bear with me here. I know that if I have an instance of MyClass called x, I can call x.GetName(). Furthermore, I could set a Func<string> variable to x.GetName. Now here's my question. Let's say I don't know the above class is called MyClass; I've got some object, x, but I have no idea what it is. I could check to see if that object has a GetName method by doing this: MethodInfo getName = x.GetType().GetMethod("GetName"); Suppose getName is not null. Then couldn't I furthermore check if getName.ReturnType == typeof(string) and getName.GetParameters().Length == 0, and at this point, wouldn't I be quite certain that the method represented by my getName object could definitely be cast to a Func<string>, somehow? I realize there's a MethodInfo.Invoke, and I also realize I could always create a Func<string> like: Func<string> getNameFunc = () => getName.Invoke(x, null); I guess what I'm asking is if there's any way to go from a MethodInfo object to the actual method it represents, incurring the performance cost of reflection in the process, but after that point being able to call the method directly (via, e.g., a Func<string> or something similar) without a performance penalty. What I'm envisioning might look something like this: // obviously this would throw an exception if GetActualInstanceMethod returned // something that couldn't be cast to a Func<string> Func<string> getNameFunc = (Func<string>)getName.GetActualInstanceMethod(x); (I realize that doesn't exist; I'm wondering if there's anything like it.) If what I'm asking doesn't make sense, or if I'm being unclear, I'll be happy to attempt to clarify.

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• Calling compiled C from R with .C()

  - by Sarah

  I'm trying to call a program (function getNBDensities in the C executable measurementDensities_out) from R. The function is passed several arrays and the variable double runsum. Right now, the getNBDensities function basically does nothing: it prints to screen the values of passed parameters. My problem is the syntax of calling the function: array(.C("getNBDensities", hr = as.double(hosp.rate), # a vector (s x 1) sp = as.double(samplingProbabilities), # another vector (s x 1) odh = as.double(odh), # another vector (s x 1) simCases = as.integer(x[c("xC1","xC2","xC3")]), # another vector (s x 1) obsCases = as.integer(y[c("yC1","yC2","yC3")]), # another vector (s x 1) runsum = as.double(runsum), # double DUP = TRUE, NAOK = TRUE, PACKAGE = "measurementDensities_out")$f, dim = length(y[c("yC1","yC2","yC3")]), dimnames = c("yC1","yC2","yC3")) The error I get, after proper execution of the function (i.e., the right output is printed to screen), is Error in dim(data) <- dim : attempt to set an attribute on NULL I'm unclear what the dimensions are that I should be passing the function: should it be s x 5 + 1 (five vectors of length s and one double)? I've tried all sorts of combinations (including sx5+1) and have found only seemingly conflicting descriptions/examples online of what's supposed to happen here. For those who are interested, the C code is below: #include <R.h> #include <Rmath.h> #include <math.h> #include <Rdefines.h> #include <R_ext/PrtUtil.h> #define NUM_STRAINS 3 #define DEBUG void getNBDensities( double *hr, double *sp, double *odh, int *simCases, int *obsCases, double *runsum ); void getNBDensities( double *hr, double *sp, double *odh, int *simCases, int *obsCases, double *runsum ) { #ifdef DEBUG for ( int s = 0; s < NUM_STRAINS; s++ ) { Rprintf("\nFor strain %d",s); Rprintf("\n\tHospitalization rate = %lg", hr[s]); Rprintf("\n\tSimulation probability = %lg",sp[s]); Rprintf("\n\tSimulated cases = %d",simCases[s]); Rprintf("\n\tObserved cases = %d",obsCases[s]); Rprintf("\n\tOverdispersion parameter = %lg",odh[s]); } Rprintf("\nRunning sum = %lg",runsum[0]); #endif } naive solution While better (i.e., potentially faster or syntactically clearer) solutions may exist (see Dirk's answer below), the following simplification of the code works: out<-.C("getNBDensities", hr = as.double(hosp.rate), sp = as.double(samplingProbabilities), odh = as.double(odh), simCases = as.integer(x[c("xC1","xC2","xC3")]), obsCases = as.integer(y[c("yC1","yC2","yC3")]), runsum = as.double(runsum)) The variables can be accessed in >out.

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• Setting up Edimax EW-7206APg as Universal Repeater

  - by Ondra Žižka

  Hi, I've troubles setting up Edimax EW-7206APg as a Universal Repeater. I've read few manuals, but they are unclear on certain points. I've managed the repeater to get to a state when it's in a "connected" state. I've set the same WPA passphrase as the router has because I haven't seen any other place to set it at. These are my settings: System Uptime 0day:1h:33m:11s Hardware Version Rev. A Runtime Code Version 1.32 Wireless Configuration Mode Universal Repeater ESSID edimax Channel Number 6 Security WPA-shared key BSSID 00:c0:9f:40:bd:38 Associated Clients 0 Wireless Repeater Interface Configuration ESSID Dusan Security WPA BSSID 00:4f:62:23:8f:7e State Connected LAN Configuration IP Address 192.168.0.10 Subnet Mask 255.255.255.0 Default Gateway 192.168.0.1 MAC Address 00:c0:9f:40:bd:37 This is ipconfig /all: Prípona DNS podle pripojení . . . : riomail.cz Popis . . . . . . . . . . . . . . : Intel(R) PRO/Wireless 2200BG Network Connection Fyzická Adresa. . . . . . . . . . : 00-0E-35-3D-77-68 Protokol DHCP povolen . . . . . . : Ano Automatická konfigurace povolena : Ano Adresa IP . . . . . . . . . . . . : 192.168.0.5 Maska podsíte . . . . . . . . . . : 255.255.255.0 Výchozí brána . . . . . . . . . . : 192.168.0.1 Server DHCP . . . . . . . . . . . : 192.168.0.1 Servery DNS . . . . . . . . . . . : 94.74.192.252 94.74.192.244 I can ping the repeater, I can ping the root AP, but not a DNS server or any other IP beyond the root AP. Anyone has an idea what's wrong? Thanks, Ondra

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• Excel: conditionally format a cell using the format of another, content-matching cell

  - by Eric A. Meyer

  I have an Excel spreadsheet where I’d like to be able to create a “key” of formatted cells with unique values, and then in another sheet format cells using the key formatting. So for example, my key is as follows, with one value per cell and the visual formatting indicated in parentheses: A (red background) B (green background) C (blue background) So that’s on one sheet (or in a remote corner of the current sheet—whichever is better). Then, in an area that I mark for conditional formatting, I can type one of those three letters and have the cell where I typed it visually formatted according to the key. So if I type a “B” into one of the conditionally formatted cells, it gets a green background. (Note that I’m using backgrounds here solely for ease of explanation: ideally I want to have all visual formatting copied over, whether it’s foreground color, background color, font weight, borders, or whatever. But I’ll take what I can get, obviously.) And—just to make it extra-tricky—if I change the formatting in the key, that change should be reflected in cells that reference the key. Thus, if I change the “B” formatting in the key from a green background to a purple background, any “B” in the main sheet should switch to the new color. Similarly, it should be possible to add or remove values from the key and have those changes applied to the main data set. I’m okay with the formatting-update-on-key-change being triggered by clicking a button or something. I suspect that if any of this is possible it will require VBA, but I’ve never used it so I’ve no idea where to start if that’s the case. I’m hoping it’s possible without VBA. I know it’s possible to just use multiple conditional formats, but my use case here is that I’m trying to create the above-described capability for someone who isn’t conversant with conditional formatting. I’d like to let them be able to define a key, update it if necessary, and keep on truckin’ without me having to rewrite the spreadsheet’s formatting rules for them. --- UPDATE --- So I think I was a bit unclear about my original request. Let me try again with an image. The image shows the “key” on the left, where values and styles are defined using keyboard and mouse input. On the right, you see the data that should be formatted to match the key. Thus if I type a “C” into a cell in the Data area, it should be blue-backed. Furthermore, if I change the formatting of “C” in the Key to have a purple background, all the “C” cells should switch from blue to purple. For further craziness, if I add more to the Key (say, “D” with a yellow background) then any “D” cells will be styled to match; if I remove a Key entry, then matching values in the Data area should revert to default styling. So. Is that more clear? Is it possible, in whole or in part? I don’t have to use conditional formatting for this; in fact, at this point I suspect I probably shouldn’t. But I’m open to any approach!

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• Set up lnux box for hosting a-z

  - by microchasm

  I am in the process of reinstalling the OS on a machine that will be used to host a couple of apps for our business. The apps will be local only; access from external clients will be via vpn only. The prior setup used a hosting control panel (Plesk) for most of the admin, and I was looking at using another similar piece of software for the reinstall - but I figured I should finally learn how it all works. I can do most of the things the software would do for me, but am unclear on the symbiosis of it all. This is all an attempt to further distance myself from the land of Configuration Programmer/Programmer, if at all possible. I can't find a full walkthrough anywhere for what I'm looking for, so I thought I'd put up this question, and if people can help me on the way I will edit this with the answers, and document my progress/pitfalls. Hopefully someday this will help someone down the line. The details: CentOS 5.5 x86_64 httpd: Apache/2.2.3 mysql: 5.0.77 (to be upgraded) php: 5.1 (to be upgraded) The requirements: SECURITY!! Secure file transfer Secure client access (SSL Certs and CA) Secure data storage Virtualhosts/multiple subdomains Local email would be nice, but not critical The Steps: Download latest CentOS DVD-iso (torrent worked great for me). Install CentOS: While going through the install, I checked the Server Components option thinking I was going to be using another Plesk-like admin. In hindsight, considering I've decided to try to go my own way, this probably wasn't the best idea. Basic config: Setup users, networking/ip address etc. Yum update/upgrade. Upgrade PHP: To upgrade PHP to the latest version, I had to look to another repo outside CentOS. IUS looks great and I'm happy I found it! cd /tmp #wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/epel-release-1-1.ius.el5.noarch.rpm #rpm -Uvh epel-release-1-1.ius.el5.noarch.rpm #wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/ius-release-1-4.ius.el5.noarch.rpm #rpm -Uvh ius-release-1-4.ius.el5.noarch.rpm yum list | grep -w \.ius\. [will list all packages available in the IUS repo] rpm -qa | grep php [will list installed packages needed to be removed. the installed packages need to be removed before you can install the IUS packages otherwise there will be conflicts] #yum shell >remove php-gd php-cli php-odbc php-mbstring php-pdo php php-xml php-common php-ldap php-mysql php-imap Setting up Remove Process >install php53 php53-mcrypt php53-mysql php53-cli php53-common php53-ldap php53-imap php53-devel >transaction solve >transaction run Leaving Shell #php -v PHP 5.3.2 (cli) (built: Apr 6 2010 18:13:45) This process removes the old version of PHP and installs the latest. To upgrade mysql: Pretty much the same process as above with PHP #/etc/init.d/mysqld stop [OK] rpm -qa | grep mysql [installed mysql packages] #yum shell >remove mysql mysql-server Setting up Remove Process >install mysql51 mysql51-server mysql51-devel >transaction solve >transaction run Leaving Shell #service mysqld start [OK] #mysql -v Server version: 5.1.42-ius Distributed by The IUS Community Project The above upgrade instructions courtesy of IUS wiki: http://wiki.iuscommunity.org/Doc/ClientUsageGuide Create a chroot jail to hold sftp user via rssh. This will force SCP/SFTP and will circumvent traditional FTP server setup. #cd /tmp #wget http://dag.wieers.com/rpm/packages/rssh/rssh-2.3.2-1.2.el5.rf.x86_64.rpm #rpm -ivh rssh-2.3.2-1.2.el5.rf.x86_64.rpm #useradd -m -d /home/dev -s /usr/bin/rssh dev #passwd dev Edit /etc/rssh.conf to grant access to SFTP to rssh users. #vi /etc/rssh.conf Uncomment line allowscp This allows me to connect to the machine via SFTP protocol in Transmit (my FTP program of choice; I'm sure it's similar with other FTP apps). Above instructions for SFTP appropriated (with appreciation!) from http://www.cyberciti.biz/tips/linux-unix-restrict-shell-access-with-rssh.html And this is where I'm at. I will keep editing this as I make progress. Any tips on how to Configure virtual interfaces/ip based virtual hosts for SSL, setting up a CA, or anything else would be appreciated.

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• moving raid 10 to another identical server both on Smart Array 6i controllers

  - by SalimQrdl

  I have dead HP DL 380G4 with RAID 1+0 with 1 logical volume from 4x72GB drives on built-in Smart Array 6i 128Mb BBWC. It was shut down properly. It seems it was usual death for Proliant with ILO led 2,3,8 lighting. I want to move array to another identical server with same raid firmware level. What is the best strategy?: I have RAID 1+0 on bay 0 bay 1 bay 2 bay 3 As I understand bay0+bay1 are in RAID 1 , bay2+bay3 are in RAID 1, and both RAID 1 pairs are in RAID 0. So should I : Clear RAID config on new server, insert bay 0, bay 2 and power-on or Create RAID 1+0 with 1 logical volume from clear HDDs , and then poweroff ,remove HDDs and insert 2 HDDs(bay 0, bay 2) from old RAID 1+0. then power-on. (each hdd has its raid position info stored but may be could work on same config) According to documentation for Smart Array 6i it could be possible to migrate. however one requirement point is unclear for me Before you move drives, the following conditions must be met: • The array is in its original configuration. " What is orginal and non-original config for RAID 1+0? Another point "If you want to move an array to another controller, you must also consider the following additional limitations: • All drives in the array must be moved at the same time." I want to move one hdd from each RAID 1 pair. to have mirrors untouched just in case. Do they mean to move all 4 simultaniously? Smart Array 6i User Guide: Moving Drives and Arrays You can move drives to other ID positionson the same array controller. You can also move a complete arrayfrom one controller to another, even if the controllers are on different servers. Before you move drives, the following conditions must be met: • If moving thedrives to a different server, the new server must have enough empty bays to accommodate all the drives simultaneously. • The move will not result in more than 14 physical drives per controller channel. • No controller will be configured with more than 32 logical volumes. • The array has no failed or missing drives. • The array is in its original configuration. • The controller is not reading from or writing to any of the spare drives in the array. • The controller is not running capacity expansion, capacity extension, or RAID or stripe size migration. • The controller is using the latestfirmware version (recommended). If you want to move an array to another controller, you must also consider the following additional limitations: • All drives in the array must be moved at the same time. • In most cases, a moved array (and the logical drives that it contains) can still undergo arraycapacity expansion, logical drive capacity extension, or migration of RAID level orstripe size. When all the conditions have been met: Back up all data before removing any drives or changing configuration. This step is requiredif you are moving data-containing drives from a controller that does not have a battery-backed cache. Power down the system. If you are moving an array from a controller that contains a RAID ADG logical volume to a controller that does not support RAID ADG: Move the drives. Power up the system. If a 1724 POST message is displayed, drive positions were changed successfully and the configuration was updated. If a 1785 (NotConfigured)POST message is displayed: a. Power down the system immediately to prevent data loss. b. Return the drives to their original locations. c. Restore the data from backup, if necessary. Check the new drive configuration byrunning ORCA or ACU ("Configuring an Array" on page 9).

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• Set up lnux box for hosting a-z [apache mysql php ssl]

  - by microchasm

  I am in the process of reinstalling the OS on a machine that will be used to host a couple of apps for our business. The apps will be local only; access from external clients will be via vpn only. The prior setup used a hosting control panel (Plesk) for most of the admin, and I was looking at using another similar piece of software for the reinstall - but I figured I should finally learn how it all works. I can do most of the things the software would do for me, but am unclear on the symbiosis of it all. This is all an attempt to further distance myself from the land of Configuration Programmer/Programmer, if at all possible. I can't find a full walkthrough anywhere for what I'm looking for, so I thought I'd put up this question, and if people can help me on the way I will edit this with the answers, and document my progress/pitfalls. Hopefully someday this will help someone down the line. The details: CentOS 5.5 x86_64 httpd: Apache/2.2.3 mysql: 5.0.77 (to be upgraded) php: 5.1 (to be upgraded) The requirements: SECURITY!! Secure file transfer Secure client access (SSL Certs and CA) Secure data storage Virtualhosts/multiple subdomains Local email would be nice, but not critical The Steps: Download latest CentOS DVD-iso (torrent worked great for me). Install CentOS: While going through the install, I checked the Server Components option thinking I was going to be using another Plesk-like admin. In hindsight, considering I've decided to try to go my own way, this probably wasn't the best idea. Basic config: Setup users, networking/ip address etc. Yum update/upgrade. Upgrade PHP: To upgrade PHP to the latest version, I had to look to another repo outside CentOS. IUS looks great and I'm happy I found it! cd /tmp #wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/epel-release-1-1.ius.el5.noarch.rpm #rpm -Uvh epel-release-1-1.ius.el5.noarch.rpm #wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/ius-release-1-4.ius.el5.noarch.rpm #rpm -Uvh ius-release-1-4.ius.el5.noarch.rpm yum list | grep -w \.ius\. [will list all packages available in the IUS repo] rpm -qa | grep php [will list installed packages needed to be removed. the installed packages need to be removed before you can install the IUS packages otherwise there will be conflicts] #yum shell >remove php-gd php-cli php-odbc php-mbstring php-pdo php php-xml php-common php-ldap php-mysql php-imap Setting up Remove Process >install php53 php53-mcrypt php53-mysql php53-cli php53-common php53-ldap php53-imap php53-devel >transaction solve >transaction run Leaving Shell #php -v PHP 5.3.2 (cli) (built: Apr 6 2010 18:13:45) This process removes the old version of PHP and installs the latest. To upgrade mysql: Pretty much the same process as above with PHP #/etc/init.d/mysqld stop [OK] rpm -qa | grep mysql [installed mysql packages] #yum shell >remove mysql mysql-server Setting up Remove Process >install mysql51 mysql51-server mysql51-devel >transaction solve >transaction run Leaving Shell #service mysqld start [OK] #mysql -v Server version: 5.1.42-ius Distributed by The IUS Community Project And this is where I'm at. I will keep editing this as I make progress. Any tips on how to Configure Virtualhosts for SSL, setting up a CA, setting up SFTP with openSSH, or anything else would be appreciated.

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• SSH new connection begins to hang (not reject or terminate) after a day or so on Ubuntu 13.04 server

  - by kross

  Recently we upgraded the server from 12.04 LTS server to 13.04. All was well, including after a reboot. With all packages updated we began to see a strange issue, ssh works for a day or so (unclear on timing) then a later request for SSH hangs (cannot ctrl+c, nothing). It is up and serving webserver traffic etc. Port 22 is open (ips etc altered slightly for posting): nmap -T4 -A x.acme.com Starting Nmap 6.40 ( http://nmap.org ) at 2013-09-12 16:01 CDT Nmap scan report for x.acme.com (69.137.56.18) Host is up (0.026s latency). rDNS record for 69.137.56.18: c-69-137-56-18.hsd1.tn.provider.net Not shown: 998 filtered ports PORT STATE SERVICE VERSION 22/tcp open ssh OpenSSH 6.1p1 Debian 4 (protocol 2.0) | ssh-hostkey: 1024 54:d3:e3:38:44:f4:20:a4:e7:42:49:d0:a7:f1:3e:21 (DSA) | 2048 dc:21:77:3b:f4:4e:74:d0:87:33:14:40:04:68:33:a6 (RSA) |_256 45:69:10:79:5a:9f:0b:f0:66:15:39:87:b9:a1:37:f7 (ECDSA) 80/tcp open http Jetty 7.6.2.v20120308 | http-title: Log in as a Bamboo user - Atlassian Bamboo |_Requested resource was http://x.acme.com/userlogin!default.action;jsessionid=19v135zn8cl1tgso28fse4d50?os_destination=%2Fstart.action Service Info: OS: Linux; CPE: cpe:/o:linux:linux_kernel Service detection performed. Please report any incorrect results at http://nmap.org/submit/ . Nmap done: 1 IP address (1 host up) scanned in 12.89 seconds Here is the ssh -vvv: ssh -vvv x.acme.com OpenSSH_5.9p1, OpenSSL 0.9.8x 10 May 2012 debug1: Reading configuration data /Users/tfergeson/.ssh/config debug1: Reading configuration data /etc/ssh_config debug1: /etc/ssh_config line 20: Applying options for * debug2: ssh_connect: needpriv 0 debug1: Connecting to x.acme.com [69.137.56.18] port 22. debug1: Connection established. debug3: Incorrect RSA1 identifier debug3: Could not load "/Users/tfergeson/.ssh/id_rsa" as a RSA1 public key debug1: identity file /Users/tfergeson/.ssh/id_rsa type 1 debug1: identity file /Users/tfergeson/.ssh/id_rsa-cert type -1 debug1: identity file /Users/tfergeson/.ssh/id_dsa type -1 debug1: identity file /Users/tfergeson/.ssh/id_dsa-cert type -1 debug1: Remote protocol version 2.0, remote software version OpenSSH_6.1p1 Debian-4 debug1: match: OpenSSH_6.1p1 Debian-4 pat OpenSSH* debug1: Enabling compatibility mode for protocol 2.0 debug1: Local version string SSH-2.0-OpenSSH_5.9 debug2: fd 3 setting O_NONBLOCK debug3: load_hostkeys: loading entries for host "x.acme.com" from file "/Users/tfergeson/.ssh/known_hosts" debug3: load_hostkeys: found key type RSA in file /Users/tfergeson/.ssh/known_hosts:10 debug3: load_hostkeys: loaded 1 keys debug3: order_hostkeyalgs: prefer hostkeyalgs: [email protected],[email protected],ssh-rsa debug1: SSH2_MSG_KEXINIT sent debug1: SSH2_MSG_KEXINIT received debug2: kex_parse_kexinit: diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: [email protected],[email protected],ssh-rsa,[email protected],[email protected],ssh-dss debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-256-96,hmac-sha2-512,hmac-sha2-512-96,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-256-96,hmac-sha2-512,hmac-sha2-512-96,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: none,[email protected],zlib debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: kex_parse_kexinit: ecdh-sha2-nistp256,ecdh-sha2-nistp384,ecdh-sha2-nistp521,diffie-hellman-group-exchange-sha256,diffie-hellman-group-exchange-sha1,diffie-hellman-group14-sha1,diffie-hellman-group1-sha1 debug2: kex_parse_kexinit: ssh-rsa,ssh-dss,ecdsa-sha2-nistp256 debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: aes128-ctr,aes192-ctr,aes256-ctr,arcfour256,arcfour128,aes128-cbc,3des-cbc,blowfish-cbc,cast128-cbc,aes192-cbc,aes256-cbc,arcfour,[email protected] debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-512,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: hmac-md5,hmac-sha1,[email protected],hmac-sha2-256,hmac-sha2-512,hmac-ripemd160,[email protected],hmac-sha1-96,hmac-md5-96 debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: none,[email protected] debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: debug2: kex_parse_kexinit: first_kex_follows 0 debug2: kex_parse_kexinit: reserved 0 debug2: mac_setup: found hmac-md5 debug1: kex: server->client aes128-ctr hmac-md5 none debug2: mac_setup: found hmac-md5 debug1: kex: client->server aes128-ctr hmac-md5 none debug1: SSH2_MSG_KEX_DH_GEX_REQUEST(1024<1024<8192) sent debug1: expecting SSH2_MSG_KEX_DH_GEX_GROUP debug2: dh_gen_key: priv key bits set: 130/256 debug2: bits set: 503/1024 debug1: SSH2_MSG_KEX_DH_GEX_INIT sent debug1: expecting SSH2_MSG_KEX_DH_GEX_REPLY debug1: Server host key: RSA dc:21:77:3b:f4:4e:74:d0:87:33:14:40:04:68:33:a6 debug3: load_hostkeys: loading entries for host "x.acme.com" from file "/Users/tfergeson/.ssh/known_hosts" debug3: load_hostkeys: found key type RSA in file /Users/tfergeson/.ssh/known_hosts:10 debug3: load_hostkeys: loaded 1 keys debug3: load_hostkeys: loading entries for host "69.137.56.18" from file "/Users/tfergeson/.ssh/known_hosts" debug3: load_hostkeys: found key type RSA in file /Users/tfergeson/.ssh/known_hosts:6 debug3: load_hostkeys: loaded 1 keys debug1: Host 'x.acme.com' is known and matches the RSA host key. debug1: Found key in /Users/tfergeson/.ssh/known_hosts:10 debug2: bits set: 493/1024 debug1: ssh_rsa_verify: signature correct debug2: kex_derive_keys debug2: set_newkeys: mode 1 debug1: SSH2_MSG_NEWKEYS sent debug1: expecting SSH2_MSG_NEWKEYS debug2: set_newkeys: mode 0 debug1: SSH2_MSG_NEWKEYS received debug1: Roaming not allowed by server debug1: SSH2_MSG_SERVICE_REQUEST sent debug2: service_accept: ssh-userauth debug1: SSH2_MSG_SERVICE_ACCEPT received debug2: key: /Users/tfergeson/.ssh/id_rsa (0x7ff189c1d7d0) debug2: key: /Users/tfergeson/.ssh/id_dsa (0x0) debug1: Authentications that can continue: publickey debug3: start over, passed a different list publickey debug3: preferred publickey,keyboard-interactive,password debug3: authmethod_lookup publickey debug3: remaining preferred: keyboard-interactive,password debug3: authmethod_is_enabled publickey debug1: Next authentication method: publickey debug1: Offering RSA public key: /Users/tfergeson/.ssh/id_rsa debug3: send_pubkey_test debug2: we sent a publickey packet, wait for reply debug1: Server accepts key: pkalg ssh-rsa blen 277 debug2: input_userauth_pk_ok: fp 3c:e5:29:6c:9d:27:d1:7d:e8:09:a2:e8:8e:6e:af:6f debug3: sign_and_send_pubkey: RSA 3c:e5:29:6c:9d:27:d1:7d:e8:09:a2:e8:8e:6e:af:6f debug1: read PEM private key done: type RSA debug1: Authentication succeeded (publickey). Authenticated to x.acme.com ([69.137.56.18]:22). debug1: channel 0: new [client-session] debug3: ssh_session2_open: channel_new: 0 debug2: channel 0: send open debug1: Requesting [email protected] debug1: Entering interactive session. debug2: callback start debug2: client_session2_setup: id 0 debug2: fd 3 setting TCP_NODELAY debug2: channel 0: request pty-req confirm 1 debug1: Sending environment. debug3: Ignored env ATLAS_OPTS debug3: Ignored env rvm_bin_path debug3: Ignored env TERM_PROGRAM debug3: Ignored env GEM_HOME debug3: Ignored env SHELL debug3: Ignored env TERM debug3: Ignored env CLICOLOR debug3: Ignored env IRBRC debug3: Ignored env TMPDIR debug3: Ignored env Apple_PubSub_Socket_Render debug3: Ignored env TERM_PROGRAM_VERSION debug3: Ignored env MY_RUBY_HOME debug3: Ignored env TERM_SESSION_ID debug3: Ignored env USER debug3: Ignored env COMMAND_MODE debug3: Ignored env rvm_path debug3: Ignored env COM_GOOGLE_CHROME_FRAMEWORK_SERVICE_PROCESS/USERS/tfergeson/LIBRARY/APPLICATION_SUPPORT/GOOGLE/CHROME_SOCKET debug3: Ignored env JPDA_ADDRESS debug3: Ignored env APDK_HOME debug3: Ignored env SSH_AUTH_SOCK debug3: Ignored env Apple_Ubiquity_Message debug3: Ignored env __CF_USER_TEXT_ENCODING debug3: Ignored env rvm_sticky_flag debug3: Ignored env MAVEN_OPTS debug3: Ignored env LSCOLORS debug3: Ignored env rvm_prefix debug3: Ignored env PATH debug3: Ignored env PWD debug3: Ignored env JAVA_HOME debug1: Sending env LANG = en_US.UTF-8 debug2: channel 0: request env confirm 0 debug3: Ignored env JPDA_TRANSPORT debug3: Ignored env rvm_version debug3: Ignored env M2_HOME debug3: Ignored env HOME debug3: Ignored env SHLVL debug3: Ignored env rvm_ruby_string debug3: Ignored env LOGNAME debug3: Ignored env M2_REPO debug3: Ignored env GEM_PATH debug3: Ignored env AWS_RDS_HOME debug3: Ignored env rvm_delete_flag debug3: Ignored env EC2_PRIVATE_KEY debug3: Ignored env RUBY_VERSION debug3: Ignored env SECURITYSESSIONID debug3: Ignored env EC2_CERT debug3: Ignored env _ debug2: channel 0: request shell confirm 1 debug2: callback done debug2: channel 0: open confirm rwindow 0 rmax 32768 I can hard reboot (only mac monitors at that location) and it will again be accessible. This now happens every single time. It is imperative that I get it sorted. The strange thing is that it behaves initially then starts to hang after several hours. I perused logs previously and nothing stood out. From the auth.log, I can see that it has allowed me in, but still I get nothing back on the client side: Sep 20 12:47:50 cbear sshd[25376]: Accepted publickey for tfergeson from 10.1.10.14 port 54631 ssh2 Sep 20 12:47:50 cbear sshd[25376]: pam_unix(sshd:session): session opened for user tfergeson by (uid=0) UPDATES: Still occurring even after setting UseDNS no and commenting out #session optional pam_mail.so standard noenv This does not appear to be a network/dns related issue, as all services running on the machine are as responsive and accessible as ever, with the exception of sshd. Any thoughts on where to start?

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• BizTalk: Internals: the Partner Direct Ports and the Orchestration Chains

  - by Leonid Ganeline

  Partner Direct Port is one of the BizTalk hidden gems. It opens simple ways to the several messaging patterns. This article based on the Kevin Lam’s blog article. The article is pretty detailed but it still leaves several unclear pieces. So I have created a sample and will show how it works from different perspectives. Requirements We should create an orchestration chain where the messages should be routed from the first stage to the second stage. The messages should not be modified. All messages has the same message type. Common artifacts Source code can be downloaded here. It is interesting but all orchestrations use only one port type. It is possible because all ports are one-way ports and use only one operation. I have added a B orchestration. It helps to test the sample, showing all test messages in channel. The Receive shape Filter is empty. A Receive Port (R_Shema1Direct) is a plain Direct Port. As you can see, a subscription expression of this direct port has only one part, the MessageType for our test schema: A Filer is empty but, as you know, a link from the Receive shape to the Port creates this MessageType expression. I use only one Physical Receive File port to send a message to all processes. Each orchestration outputs a Trace.WriteLine(“<Orchestration Name>”). Forward Binding This sample has three orchestrations: A_1, A_21 and A_22. A_1 is a sender, A_21 and A_22 are receivers. Here is a subscription of the A_1 orchestration: It has two parts A MessageType. The same was for the B orchestration. A ReceivePortID. There was no such parameter for the B orchestration. It was created because I have bound the orchestration port with Physical Receive File port. This binding means the PortID parameter is added to the subscription. How to set up the ports? All ports involved in the message exchange should be the same port type. It forces us to use the same operation and the same message type for the bound ports. This step as absolutely contra-intuitive. We have to choose a Partner Orchestration parameter for the sending orchestration, A_1. The first strange thing is it is not a partner orchestration we have to choose but an orchestration port. But the most strange thing is we have to choose exactly this orchestration and exactly this port.It is not a port from the partner, receive orchestrations, A_21 or A_22, but it is A_1 orchestration and S_SentFromA_1 port. Now we have to choose a Partner Orchestration parameter for the received orchestrations, A_21 and A_22. Nothing strange is here except a parameter name. We choose the port of the sender, A_1 orchestration and S_SentFromA_1 port. As you can see the Partner Orchestration parameter for the sender and receiver orchestrations is the same. Testing I dropped a test file in a file folder. There we go: A dropped file was received by B and by A_1 A_1 sent a message forward. A message was received by B, A_21, A_22 Let’s look at a context of a message sent by A_1 on the second step: A MessageType part. It is quite expected. A PartnerService, a ParnerPort, an Operation. All those parameters were set up in the Partner Orchestration parameter on both bound ports.     Now let’s see a subscription of the A_21 and A_22 orchestrations. Now it makes sense. That’s why we have chosen such a strange value for the Partner Orchestration parameter of the sending orchestration. Inverse Binding This sample has three orchestrations: A_11, A_12 and A_2. A_11 and A_12 are senders, A_2 is receiver. How to set up the ports? All ports involved in the message exchange should be the same port type. It forces us to use the same operation and the same message type for the bound ports. This step as absolutely contra-intuitive. We have to choose a Partner Orchestration parameter for a receiving orchestration, A_2. The first strange thing is it is not a partner orchestration we have to choose but an orchestration port. But the most strange thing is we have to choose exactly this orchestration and exactly this port.It is not a port from the partner, sent orchestrations, A_11 or A_12, but it is A_2 orchestration and R_SentToA_2 port. Now we have to choose a Partner Orchestration parameter for the sending orchestrations, A_11 and A_12. Nothing strange is here except a parameter name. We choose the port of the sender, A_2 orchestration and R_SentToA_2 port. Testing I dropped a test file in a file folder. There we go: A dropped file was received by B, A_11 and by A_12 A_11 and A_12 sent two messages forward. The messages were received by B, A_2 Let’s see what was a context of a message sent by A_1 on the second step: A MessageType part. It is quite expected. A PartnerService, a ParnerPort, an Operation. All those parameters were set up in the Partner Orchestration parameter on both bound ports. Here is a subscription of the A_2 orchestration. Models I had a hard time trying to explain the Partner Direct Ports in simple terms. I have finished with this model: Forward Binding Receivers know a Sender. Sender doesn’t know Receivers. Publishers know a Subscriber. Subscriber doesn’t know Publishers. 1 –> 1 1 –> M Inverse Binding Senders know a Receiver. Receiver doesn’t know Senders. Subscribers know a Publisher. Publisher doesn’t know Subscribers. 1 –> 1 M –> 1 Notes   Orchestration chain It’s worth to note, the Partner Direct Port Binding creates a chain opened from one side and closed from another. The Forward Binding: A new Receiver can be added at run-time. The Sender can not be changed without design-time changes in Receivers. The Inverse Binding: A new Sender can be added at run-time. The Receiver can not be changed without design-time changes in Senders.

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• How to begin? Windows 8 Development

  - by Dennis Vroegop

  Ok. I convinced you in my last post to do some Win8 development. You want a piece of that cake, or whatever your reasons may be. Good! Welcome to the club! Now let me ask you a question: what are you going to write? Ah. That’s the big one, isn’t it? What indeed? If you have been creating applications for computers before you’re in for quite a shock. The way people perceive apps on a tablet is quite different from what we know as applications. There’s a reason we call them apps instead of applications! Yes, technically they are applications but we don’t call them apps only because it sounds cool. The abbreviated form of the word applications itself is a pointer. Apps are small. Apps are focused. Apps are more lightweight. Apps do one thing but they do that one thing extremely good. In the ‘old’ days we wrote huge systems. We build ecosystems of services, screens, databases and more to create a system that provides value for the user. Think about it: what application do you use most at work? Can you in one sentence describe what it is, or what it does and yet still distinctively describe its purpose? I doubt you can. Let’s have a look at Outlouk. We all know it and we all love or hate it. But what is it? A mail program? No, there’s so much more there: calendar, contacts, RSS feeds and so on. Some call it a ‘collaboration’  application but that’s not really true as well. After all, why should a collaboration application give me my schedule for the day? I think the best way to describe Outlook is “client for Exchange”  although that isn’t accurate either. Anyway: Outlook is a great application but it’s not an ‘app’ and therefor not very suitable for WinRT. Ok. Disclaimer here: yes, you can write big applications for WinRT. Some will. But that’s not what 99.9% of the developers will do. So I am stating here that big applications are not meant for WinRT. If 0.01% of the developers think that this is nonsense then they are welcome to go ahead but for the majority here this is not what we’re talking about. So: Apps are small, lightweight and good at what they do but only at that. If you’re a Phone developer you already know that: Phone apps on any platform fit the description I have above. If you’ve ever worked in a large cooperation before you might have seen one of these before: the Mission Statement. It’s supposed to be a oneliner that sums up what the company is supposed to do. Funny enough: although this doesn’t work for large companies it does work for defining your app. A mission statement for an app describes what it does. If it doesn’t fit in the mission statement then your app is going to get to big and will fail. A statement like this should be in the following style “<your app name> is the best app to <describe single task>” Fill in the blanks, write it and go! Mmm.. not really. There are some things there we need to think about. But the statement is a very, very important one. If you cannot fit your app in that line you’re preparing to fail. Your app will become to big, its purpose will be unclear and it will be hard to use. People won’t download it and those who do will give it a bad rating therefor preventing that huge success you’ve been dreaming about. Stick to the statement! Ok, let’s give it a try: “PlanesAreCool” is the best app to do planespotting in the field. You might have seen these people along runways of airports: taking photographs of airplanes and noting down their numbers and arrival- and departure times. We are going to help them out with our great app! If you look at the statement, can you guess what it does? I bet you can. If you find out it isn’t clear enough of if it’s too broad, refine it. This is probably the most important step in the development of your app so give it enough time! So. We’ve got the statement. Print it out, stick it to the wall and look at it. What does it tell you? If you see this, what do you think the app does? Write that down. Sit down with some friends and talk about it. What do they expect from an app like this? Write that down as well. Brainstorm. Make a list of features. This is mine: Note planes Look up aircraft carriers Add pictures of that plane Look up airfields Notify friends of new spots Look up details of a type of plane Plot a graph with arrival and departure times Share new spots on social media Look up history of a particular aircraft Compare your spots with friends Write down arrival times Write down departure times Write down wind conditions Write down the runway they take Look up weather conditions for next spotting day Invite friends to join you for a day of spotting. Now, I must make it clear that I am not a planespotter nor do I know what one does. So if the above list makes no sense, I apologize. There is a lesson: write apps for stuff you know about…. First of all, let’s look at our statement and then go through the list of features. Remove everything that has nothing to do with that statement! If you end up with an empty list, try again with both steps. Note planes Look up aircraft carriers Add pictures of that plane Look up airfields Notify friends of new spots Look up details of a type of plane Plot a graph with arrival and departure times Share new spots on social media Look up history of a particular aircraft Compare your spots with friends Write down arrival times Write down departure times Write down wind conditions Write down the runway they take Look up weather conditions for next spotting day Invite friends to join you for a day of spotting. That's better. The things I removed could be pretty useful to a plane spotter and could be fun to write. But do they match the statement? I said that the app is for spotting in the field, so “look up airfields” doesn’t belong there: I know where I am so why look it up? And the same goes for inviting friends or looking up the weather conditions for tomorrow. I am at the airfield right now, looking through my binoculars at the planes. I know the weather now and I don’t care about tomorrow. If you feel the items you’ve crossed out are valuable, then why not write another app? One that says “SpotNoter” is the best app for preparing a day of spotting with my friends. That’s a different app! Remember: Win8 apps are small and very good at doing ONE thing, and one thing only! If you have made that list, it’s time to prepare the navigation of your app. The navigation is how users see your app and how they use it. We’ll do that next time!

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• Standards Corner: OAuth WG Client Registration Problem

  - by Tanu Sood

  Phil Hunt is an active member of multiple industry standards groups and committees (see brief bio at the end of the post) and has spearheaded discussions, creation and ratifications of  Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-family:"Calibri","sans-serif"; mso-ascii- mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi- mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} industry standards including the Kantara Identity Governance Framework, among others. Being an active voice in the industry standards development world, we have invited him to share his discussions, thoughts, news & updates, and discuss use cases, implementation success stories (and even failures) around industry standards on this monthly column. Author: Phil Hunt This afternoon, the OAuth Working Group will meet at IETF88 in Vancouver to discuss some important topics important to the maturation of OAuth. One of them is the OAuth client registration problem.OAuth (RFC6749) was initially developed with a simple deployment model where there is only monopoly or singleton cloud instance of a web API (e.g. there is one Facebook, one Google, on LinkedIn, and so on). When the API publisher and API deployer are the same monolithic entity, it easy for developers to contact the provider and register their app to obtain a client_id and credential.But what happens when the API is for an open source project where there may be 1000s of deployed copies of the API (e.g. such as wordpress). In these cases, the authors of the API are not the people running the API. In these scenarios, how does the developer obtain a client_id? An example of an "open deployed" API is OpenID Connect. Connect defines an OAuth protected resource API that can provide personal information about an authenticated user -- in effect creating a potentially common API for potential identity providers like Facebook, Google, Microsoft, Salesforce, or Oracle. In Oracle's case, Fusion applications will soon have RESTful APIs that are deployed in many different ways in many different environments. How will developers write apps that can work against an openly deployed API with whom the developer can have no prior relationship?At present, the OAuth Working Group has two proposals two consider: Dynamic RegistrationDynamic Registration was originally developed for OpenID Connect and UMA. It defines a RESTful API in which a prospective client application with no client_id creates a new client registration record with a service provider and is issued a client_id and credential along with a registration token that can be used to update registration over time.As proof of success, the OIDC community has done substantial implementation of this spec and feels committed to its use. Why not approve?Well, the answer is that some of us had some concerns, namely: Recognizing instances of software - dynamic registration treats all clients as unique. It has no defined way to recognize that multiple copies of the same client are being registered other then assuming if the registration parameters are similar it might be the same client. Versioning and Policy Approval of open APIs and clients - many service providers have to worry about change management. They expect to have approval cycles that approve versions of server and client software for use in their environment. In some cases approval might be wide open, but in many cases, approval might be down to the specific class of software and version. Registration updates - when does a client actually need to update its registration? Shouldn't it be never? Is there some characteristic of deployed code that would cause it to change? Options lead to complexity - because each client is treated as unique, it becomes unclear how the clients and servers will agree on what credentials forms are acceptable and what OAuth features are allowed and disallowed. Yet the reality is, developers will write their application to work in a limited number of ways. They can't implement all the permutations and combinations that potential service providers might choose. Stateful registration - if the primary motivation for registration is to obtain a client_id and credential, why can't this be done in a stateless fashion using assertions? Denial of service - With so much stateful registration and the need for multiple tokens to be issued, will this not lead to a denial of service attack / risk of resource depletion? At the very least, because of the information gathered, it would difficult for service providers to clean up "failed" registrations and determine active from inactive or false clients. There has yet to be much wide-scale "production" use of dynamic registration other than in small closed communities. Client Association A second proposal, Client Association, has been put forward by Tony Nadalin of Microsoft and myself. We took at look at existing use patterns to come up with a new proposal. At the Berlin meeting, we considered how WS-STS systems work. More recently, I took a review of how mobile messaging clients work. I looked at how Apple, Google, and Microsoft each handle registration with APNS, GCM, and WNS, and a similar pattern emerges. This pattern is to use an existing credential (mutual TLS auth), or client bearer assertion and swap for a device specific bearer assertion.In the client association proposal, the developer's registration with the API publisher is handled by having the developer register with an API publisher (as opposed to the party deploying the API) and obtaining a software "statement". Or, if there is no "publisher" that can sign a statement, the developer may include their own self-asserted software statement.A software statement is a special type of assertion that serves to lock application registration profile information in a signed assertion. The statement is included with the client application and can then be used by the client to swap for an instance specific client assertion as defined by section 4.2 of the OAuth Assertion draft and profiled in the Client Association draft. The software statement provides a way for service provider to recognize and configure policy to approve classes of software clients, and simplifies the actual registration to a simple assertion swap. Because the registration is an assertion swap, registration is no longer "stateful" - meaning the service provider does not need to store any information to support the client (unless it wants to). Has this been implemented yet? Not directly. We've only delivered draft 00 as an alternate way of solving the problem using well-known patterns whose security characteristics and scale characteristics are well understood. Dynamic Take II At roughly the same time that Client Association and Software Statement were published, the authors of Dynamic Registration published a "split" version of the Dynamic Registration (draft-richer-oauth-dyn-reg-core and draft-richer-oauth-dyn-reg-management). While some of the concerns above are addressed, some differences remain. Registration is now a simple POST request. However it defines a new method for issuing client tokens where as Client Association uses RFC6749's existing extension point. The concern here is whether future client access token formats would be addressed properly. Finally, Dyn-reg-core does not yet support software statements. Conclusion The WG has some interesting discussion to bring this back to a single set of specifications. Dynamic Registration has significant implementation, but Client Association could be a much improved way to simplify implementation of the overall OpenID Connect specification and improve adoption. In fairness, the existing editors have already come a long way. Yet there are those with significant investment in the current draft. There are many that have expressed they don't care. They just want a standard. There is lots of pressure on the working group to reach consensus quickly.And that folks is how the sausage is made.Note: John Bradley and Justin Richer recently published draft-bradley-stateless-oauth-client-00 which on first look are getting closer. Some of the details seem less well defined, but the same could be said of client-assoc and software-statement. I hope we can merge these specs this week. Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-family:"Calibri","sans-serif"; mso-ascii- mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi- mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;} About the Writer: Phil Hunt joined Oracle as part of the November 2005 acquisition of OctetString Inc. where he headed software development for what is now Oracle Virtual Directory. Since joining Oracle, Phil works as CMTS in the Identity Standards group at Oracle where he developed the Kantara Identity Governance Framework and provided significant input to JSR 351. Phil participates in several standards development organizations such as IETF and OASIS working on federation, authorization (OAuth), and provisioning (SCIM) standards.  Phil blogs at www.independentid.com and a Twitter handle of @independentid.

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• The future for Microsoft

  - by Scott Dorman

  Originally posted on: http://geekswithblogs.net/sdorman/archive/2013/10/16/the-future-for-microsoft.aspxMicrosoft is in the process of reinventing itself. While some may argue that it’s “too little, too late” or that their growing consumer-focused strategy is wrong, the truth of the situation is that Microsoft is reinventing itself into a new company. While Microsoft is now calling themselves a “devices and services” company, that’s not entirely accurate. Let’s look at some facts: Microsoft will always (for the long-term foreseeable future) be financially split into the following divisions: Windows/Operating Systems, which for FY13 made up approximately 24% of overall revenue. Server and Tools, which for FY13 made up approximately 26% of overall revenue. Enterprise/Business Products, which for FY13 made up approximately 32% of overall revenue. Entertainment and Devices, which for FY13 made up approximately 13% of overall revenue. Online Services, which for FY13 made up approximately 4% of overall revenue. It is important to realize that hardware products like the Surface fall under the Windows/Operating Systems division while products like the Xbox 360 fall under the Entertainment and Devices division. (Presumably other hardware, such as mice, keyboards, and cameras, also fall under the Entertainment and Devices division.) It’s also unclear where Microsoft’s recent acquisition of Nokia’s handset division will fall, but let’s assume that it will be under Entertainment and Devices as well. Now, for the sake of argument, let’s assume a slightly different structure that I think is more in line with how Microsoft presents itself and how the general public sees it: Consumer Products and Devices, which would probably make up approximately 9% of overall revenue. Developer Tools, which would probably make up approximately 13% of overall revenue. Enterprise Products and Devices, which would probably make up approximately 47% of overall revenue. Entertainment, which would probably make up approximately 13% of overall revenue. Online Services, which would probably make up approximately 17% of overall revenue. (Just so we’re clear, in this structure hardware products like the Surface, a portion of Windows sales, and other hardware fall under the Consumer Products and Devices division. I’m assuming that more of the income for the Windows division is coming from enterprise/volume licenses so 15% of that income went to the Enterprise Products and Devices division. Most of the enterprise services, like Azure, fall under the Online Services division so half of the Server and Tools income went there as well.) No matter how you look at it, the bulk of Microsoft’s income still comes from not just the enterprise but also software sales, and this really shouldn’t surprise anyone. So, now that the stage is set…what’s the future for Microsoft? The future I see for Microsoft (again, this is just my prediction based on my own instinct, gut-feel and publicly available information) is this: Microsoft is becoming a consumer-focused enterprise company. Let’s look at it a different way. Microsoft is an enterprise-focused company trying to create a larger consumer presence.  To a large extent, this is the exact opposite of Apple, who is really a consumer-focused company trying to create a larger enterprise presence. The major reason consumer-focused companies (like Apple) have started making in-roads into the enterprise is the “bring your own device” phenomenon. Yes, Apple has created some “game-changing” products but their enterprise influence is still relatively small. Unfortunately (for this blog post at least), Apple provides revenue in terms of hardware products rather than business divisions, so it’s not possible to do a direct comparison. However, in the interest of transparency, from Apple’s Quarterly Report (filed 24 July 2013), their revenue breakdown is: iPhone, which for the 3 months ending 29 June 2013 made up approximately 51% of revenue. iPad, which for the 3 months ending 29 June 2013 made up approximately 18% of revenue. Mac, which for the 3 months ending 29 June 2013 made up approximately 14% of revenue. iPod, which for the 3 months ending 29 June 2013 made up approximately 2% of revenue. iTunes, Software, and Services, which for the 3 months ending 29 June 2013 made up approximately 11% of revenue. Accessories, which for the 3 months ending 29 July 2013 made up approximately 3% of revenue. From this, it’s pretty clear that Apple is a consumer-and-hardware-focused company. At this point, you may be asking yourself “Where is all of this going?” The answer to that lies in Microsoft’s shift in company focus. They are becoming more consumer focused, but what exactly does that mean? The biggest change (at least that’s been in the news lately) is the pending purchase of Nokia’s handset division. This, in combination with their Surface line of tablets and the Xbox, will put Microsoft squarely in the realm of a hardware-focused company in addition to being a software-focused company. That can (and most likely will) shift the revenue split to looking at revenue based on software sales (both consumer and enterprise) and also hardware sales (mostly on the consumer side). If we look at things strictly from a Windows perspective, Microsoft clearly has a lot of irons in the fire at the moment. Discounting the various product SKUs available and painting the picture with broader strokes, there are currently 5 different Windows-based operating systems: Windows Phone Windows Phone 7.x, which runs on top of the Windows CE kernel Windows Phone 8.x+, which runs on top of the Windows 8 kernel Windows RT The ARM-based version of Windows 8, which runs on top of the Windows 8 kernel Windows (Pro) The Intel-based version of Windows 8, which runs on top of the Windows 8 kernel Xbox The Xbox 360, which runs it’s own proprietary OS. The Xbox One, which runs it’s own proprietary OS, a version of Windows running on top of the Windows 8 kernel and a proprietary “manager” OS which manages the other two. Over time, Windows Phone 7.x devices will fade so that really leaves 4 different versions. Looking at Windows RT and Windows Phone 8.x paints an interesting story. Right now, all mobile phone devices run on some sort of ARM chip and that doesn’t look like it will change any time soon. That means Microsoft has two different Windows based operating systems for the ARM platform. Long term, it doesn’t make sense for Microsoft to continue supporting that arrangement. I have long suspected (since the Surface was first announced) that Microsoft will unify these two variants of Windows and recent speculation from some of the leading Microsoft watchers lends credence to this suspicion. It is rumored that upcoming Windows Phone releases will include support for larger screen sizes, relax the requirement to have a hardware-based back button and will continue to improve API parity between Windows Phone and Windows RT. At the same time, Windows RT will include support for smaller screen sizes. Since both of these operating systems are based on the same core Windows kernel, it makes sense (both from a financial and development resource perspective) for Microsoft to unify them. The user interfaces are already very similar. So similar in fact, that visually it’s difficult to tell them apart. To illustrate this, here are two screen captures: Other than a few variations (the Bing News app, the picture shown in the Pictures tile and the spacing between the tiles) these are identical. The one on the left is from my Windows 8.1 laptop (which looks the same as on my Surface RT) and the one on the right is from my Windows Phone 8 Lumia 925. This pretty clearly shows that from a consumer perspective, there really is no practical difference between how these two operating systems look and how you interact with them. For the consumer, your entertainment device (Xbox One), phone (Windows Phone) and mobile computing device (Surface [or some other vendors tablet], laptop, netbook or ultrabook) and your desktop computing device (desktop) will all look and feel the same. While many people will denounce this consistency of user experience, I think this will be a good thing in the long term, especially for the upcoming generations. For example, my 5-year old son knows how to use my tablet, phone and Xbox because they all feature nearly identical user experiences. When Windows 8 was released, Microsoft allowed a Windows Store app to be purchased once and installed on as many as 5 devices. With Windows 8.1, this limit has been increased to over 50. Why is that important? If you consider that your phone, computing devices, and entertainment device will be running the same operating system (with minor differences related to physical hardware chipset), that means that I could potentially purchase my sons favorite Angry Birds game once and be able to install it on all of the devices I own. (And for those of you wondering, it’s only 7 [at the moment].) From an app developer perspective, the story becomes even more compelling. Right now there are differences between the different operating systems, but those differences are shrinking. The user interface technology for both is XAML but there are different controls available and different user experience concepts. Some of the APIs available are the same while some are not. You can’t develop a Windows Phone app that can also run on Windows (either Windows Pro or RT). With each release of Windows Phone and Windows RT, those difference become smaller and smaller. Add to this mix the Xbox One, which will also feature a Windows-based operating system and the same “modern” (tile-based) user interface and the visible distinctions between the operating systems will become even smaller. Unifying the operating systems means one set of APIs and one code base to maintain for an app that can run on multiple devices. One code base means it’s easier to add features and fix bugs and that those changes become available on all devices at the same time. It also means a single app store, which will increase the discoverability and reach of your app and consolidate revenue and app profile management. Now, the choice of what devices an app is available on becomes a simple checkbox decision rather than a technical limitation. Ultimately, this means more apps available to consumers, which is always good for the app ecosystem. Is all of this just rumor, speculation and conjecture? Of course, but it’s not unfounded. As I mentioned earlier, some of the prominent Microsoft watchers are also reporting similar rumors. However, Microsoft itself has even hinted at this future with their recent organizational changes and by telling developers “if you want to develop for Xbox One, start developing for Windows 8 now.” I think this pretty clearly paints the following picture: Microsoft is committed to the “modern” user interface paradigm. Microsoft is changing their release cadence (for all products, not just operating systems) to be faster and more modular. Microsoft is going to continue to unify their OS platforms both from a consumer perspective and a developer perspective. While this direction will certainly concern some people it will excite many others. Microsoft’s biggest failing has always been following through with a strong and sustained marketing strategy that presents a consistent view point and highlights what this unified and connected experience looks like and how it benefits consumers and enterprises. We’ve started to see some of this over the last few years, but it needs to continue and become more aggressive and consistent. In the long run, I think Microsoft will be able to pull all of these technologies and devices together into one seamless ecosystem. It isn’t going to happen overnight, but my prediction is that we will be there by the end of 2016. As both a consumer and a developer, I, for one, am excited about the future of Microsoft.

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• Set up linux box for hosting a-z

  - by microchasm

  I am in the process of reinstalling the OS on a machine that will be used to host a couple of apps for our business. The apps will be local only; access from external clients will be via vpn only. The prior setup used a hosting control panel (Plesk) for most of the admin, and I was looking at using another similar piece of software for the reinstall - but I figured I should finally learn how it all works. I can do most of the things the software would do for me, but am unclear on the symbiosis of it all. This is all an attempt to further distance myself from the land of Configuration Programmer/Programmer, if at all possible. I can't find a full walkthrough anywhere for what I'm looking for, so I thought I'd put up this question, and if people can help me on the way I will edit this with the answers, and document my progress/pitfalls. Hopefully someday this will help someone down the line. The details: CentOS 5.5 x86_64 httpd: Apache/2.2.3 mysql: 5.0.77 (to be upgraded) php: 5.1 (to be upgraded) The requirements: SECURITY!! Secure file transfer Secure client access (SSL Certs and CA) Secure data storage Virtualhosts/multiple subdomains Local email would be nice, but not critical The Steps: Download latest CentOS DVD-iso (torrent worked great for me). Install CentOS: While going through the install, I checked the Server Components option thinking I was going to be using another Plesk-like admin. In hindsight, considering I've decided to try to go my own way, this probably wasn't the best idea. Basic config: Setup users, networking/ip address etc. Yum update/upgrade. Upgrade PHP/MySQL: To upgrade PHP and MySQL to the latest versions, I had to look to another repo outside CentOS. IUS looks great and I'm happy I found it! Add IUS repository to our package manager cd /tmp wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/epel-release-1-1.ius.el5.noarch.rpm rpm -Uvh epel-release-1-1.ius.el5.noarch.rpm wget http://dl.iuscommunity.org/pub/ius/stable/Redhat/5/x86_64/ius-release-1-4.ius.el5.noarch.rpm rpm -Uvh ius-release-1-4.ius.el5.noarch.rpm yum list | grep -w \.ius\. # list all the packages in the IUS repository; use this to find PHP/MySQL version and libraries you want to install Remove old version of PHP and install newer version from IUS rpm -qa | grep php # to list all of the installed php packages we want to remove yum shell # open an interactive yum shell remove php-common php-mysql php-cli #remove installed PHP components install php53 php53-mysql php53-cli php53-common #add packages you want transaction solve #important!! checks for dependencies transaction run #important!! does the actual installation of packages. [control+d] #exit yum shell php -v PHP 5.3.2 (cli) (built: Apr 6 2010 18:13:45) Upgrade MySQL from IUS repository /etc/init.d/mysqld stop rpm -qa | grep mysql # to see installed mysql packages yum shell remove mysql mysql-server #remove installed MySQL components install mysql51 mysql51-server mysql51-devel transaction solve #important!! checks for dependencies transaction run #important!! does the actual installation of packages. [control+d] #exit yum shell service mysqld start mysql -v Server version: 5.1.42-ius Distributed by The IUS Community Project Upgrade instructions courtesy of IUS wiki: http://wiki.iuscommunity.org/Doc/ClientUsageGuide Install rssh (restricted shell) to provide scp and sftp access, without allowing ssh login cd /tmp wget http://dag.wieers.com/rpm/packages/rssh/rssh-2.3.2-1.2.el5.rf.x86_64.rpm rpm -ivh rssh-2.3.2-1.2.el5.rf.x86_64.rpm useradd -m -d /home/dev -s /usr/bin/rssh dev passwd dev Edit /etc/rssh.conf to grant access to SFTP to rssh users. vi /etc/rssh.conf Uncomment or add: allowscp allowsftp This allows me to connect to the machine via SFTP protocol in Transmit (my FTP program of choice; I'm sure it's similar with other FTP apps). rssh instructions appropriated (with appreciation!) from http://www.cyberciti.biz/tips/linux-unix-restrict-shell-access-with-rssh.html Set up virtual interfaces ifconfig eth1:1 192.168.1.3 up #start up the virtual interface cd /etc/sysconfig/network-scripts/ cp ifcfg-eth1 ifcfg-eth1:1 #copy default script and match name to our virtual interface vi ifcfg-eth1:1 #modify eth1:1 script #ifcfg-eth1:1 | modify so it looks like this: DEVICE=eth1:1 IPADDR=192.168.1.3 NETMASK=255.255.255.0 NETWORK=192.168.1.0 ONBOOT=yes NAME=eth1:1 Add more Virtual interfaces as needed by repeating. Because of the ONBOOT=yes line in the ifcfg-eth1:1 file, this interface will be brought up when the system boots, or the network starts/restarts. service network restart Shutting down interface eth0: [ OK ] Shutting down interface eth1: [ OK ] Shutting down loopback interface: [ OK ] Bringing up loopback interface: [ OK ] Bringing up interface eth0: [ OK ] Bringing up interface eth1: [ OK ] ping 192.168.1.3 64 bytes from 192.168.1.3: icmp_seq=1 ttl=64 time=0.105 ms And this is where I'm at. I will keep editing this as I make progress. Any tips on how to Configure virtual interfaces/ip based virtual hosts for SSL, setting up a CA, or anything else would be appreciated.

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• Navigation in Win8 Metro Style applications

  - by Dennis Vroegop

  In Windows 8, Touch is, as they say, a first class citizen. Now, to be honest: they also said that in Windows 7. However in Win8 this is actually true. Applications are meant to be used by touch. Yes, you can still use mouse, keyboard and pen and your apps should take that into account but touch is where you should focus on initially. Will all users have touch enabled devices? No, not in the first place. I don’t think touchscreens will be on every device sold next year. But in 5 years? Who knows? Don’t forget: if your app is successful it will be around for a long time and by that time touchscreens will be everywhere. Another reason to embrace touch is that it’s easier to develop a touch-oriented app and then to make sure that keyboard, nouse and pen work as doing it the other way around. Porting a mouse-based application to a touch based application almost never works. The reverse gives you much more chances for success. That being said, there are some things that you need to think about. Most people have more than one finger, while most users only use one mouse at the time. Still, most touch-developers translate their mouse-knowledge to the touch and think they did a good job. Martin Tirion from Microsoft said that since Touch is a new language people face the same challenges they do when learning a new real spoken language. The first thing people try when learning a new language is simply replace the words in their native language to the newly learned words. At first they don’t care about grammar. To a native speaker of that other language this sounds all wrong but they still will be able to understand what the intention was. If you don’t believe me: try Google translate to translate something for you from your language to another and then back and see what happens. The same thing happens with Touch. Most developers translate a mouse-click into a tap-event and think they’re done. Well matey, you’re not done. Not by far. There are things you can do with a mouse that you cannot do with touch. Think hover. A mouse has the ability to ‘slide’ over UI elements. Touch doesn’t (I know: with Pen you can do this but I’m talking about actual fingers here). A touch is either there or it isn’t. And right-click? Forget about it. A click is a click.  Yes, you have more than one finger but the machine doesn’t know which finger you use… The other way around is also true. Like I said: most users only have one mouse but they are likely to have more than one finger. So how do we take that into account? Thinking about this is really worth the time: you might come up with some surprisingly good ideas! Still: don’t forget that not every user has touch-enabled hardware so make sure your app is useable for both groups. Keep this in mind: we’re going to need it later on! Now. Apps should be easy to use. You don’t want your user to read through pages and pages of documentation before they can use the app. Imagine that spotter next to an airfield suddenly seeing a prototype of a Concorde 2 landing on the nearby runway. He probably wants to enter that information in our app NOW and not after he’s taken a 3 day course. Even if he still has to download the app, install it for the first time and then run it he should be on his way immediately. At least, fast enough to note down the details of that unique, rare and possibly exciting sighting he just did. So.. How do we do this? Well, I am not talking about games here. Games are in a league of their own. They fall outside the scope of the apps I am describing. But all the others can roughly be characterized as being one of two flavors: the navigation is either flat or hierarchical. That’s it. And if it’s hierarchical it’s no more than three levels deep. Not more. Your users will get lost otherwise and we don’t want that. Flat is simple. Just imagine we have one screen that is as high as our physical screen is and as wide as you need it to be. Don’t worry if it doesn’t fit on the screen: people can scroll to the right and left. Don’t combine up/down and left/right scrolling: it’s confusing. Next to that, since most users will hold their device in landscape mode it’s very natural to scroll horizontal. So let’s use that when we have a flat model. The same applies to the hierarchical model. Try to have at most three levels. If you need more space, find a way to group the items in such a way that you can fit it in three, very wide lanes. At the highest level we have the so called hub level. This is the entry point of the app and as such it should give the user an immediate feeling of what the app is all about. If your app has categories if items then you might show these categories here. And while you’re at it: also show 2 or 3 of the items itself here to give the user a taste of what lies beneath. If the user selects a category you go to the section part. Here you show several sections (again, go as wide as you need) with again some detail examples. After that: the details layer shows each item. By giving some samples of the underlaying layer you achieve several things: you make the layer attractive by showing several different things, you show some highlights so the user sees actual content and you provide a shortcut to the layers underneath. The image below is borrowed from the http://design.windows.com website which has tons and tons of examples: For our app we’ll use this layout. So what will we show? Well, let’s see what sorts of features our app has to offer. I’ll repeat them here: Note planes Add pictures of that plane Notify friends of new spots Share new spots on social media Write down arrival times Write down departure times Write down the runway they take I am sure you can think of some more items but for now we'll use these. In the hub we’ll show something that represents “Spots”, “Friends”, “Social”. Apparently we have an inner list of spotter-friends that are in the app, while we also have to whole world in social. In the layer below we show something else, depending on what the user choose. When they choose “Spots” we’ll display the last spots, last spots by our friends (so we can actually jump from this category to the one next to it) and so on. When they choose a “spot” (or press the + icon in the App bar, which I’ll talk about next time) they go to the lowest and final level that shows details about that spot, including a picture, date and time and the notes belonging to that entry. You’d be amazed at how easy it is to organize your app this way. If you don’t have enough room in these three layers you probably could easily get away with grouping items. Take a look at our hub: we have three completely different things in one place. If you still can’t fit it all in in a logical and consistent way, chances are you are trying to do too much in this app. Go back to your mission statement, determine if it is specific enough and if your feature list helps that statement or makes it unclear. Go ahead. Give it a go! Next time we’ll talk about the look and feel, the charms and the app-bar….

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• Linker error when compiling boost.asio example

  - by Alon

  Hi, I'm trying to learn a little bit C++ and Boost.Asio. I'm trying to compile the following code example: #include <iostream> #include <boost/array.hpp> #include <boost/asio.hpp> using boost::asio::ip::tcp; int main(int argc, char* argv[]) { try { if (argc != 2) { std::cerr << "Usage: client <host>" << std::endl; return 1; } boost::asio::io_service io_service; tcp::resolver resolver(io_service); tcp::resolver::query query(argv[1], "daytime"); tcp::resolver::iterator endpoint_iterator = resolver.resolve(query); tcp::resolver::iterator end; tcp::socket socket(io_service); boost::system::error_code error = boost::asio::error::host_not_found; while (error && endpoint_iterator != end) { socket.close(); socket.connect(*endpoint_iterator++, error); } if (error) throw boost::system::system_error(error); for (;;) { boost::array<char, 128> buf; boost::system::error_code error; size_t len = socket.read_some(boost::asio::buffer(buf), error); if (error == boost::asio::error::eof) break; // Connection closed cleanly by peer. else if (error) throw boost::system::system_error(error); // Some other error. std::cout.write(buf.data(), len); } } catch (std::exception& e) { std::cerr << e.what() << std::endl; } return 0; } With the following command line: g++ -I /usr/local/boost_1_42_0 a.cpp and it throws an unclear error: /tmp/ccCv9ZJA.o: In function `__static_initialization_and_destruction_0(int, int)': a.cpp:(.text+0x654): undefined reference to `boost::system::get_system_category()' a.cpp:(.text+0x65e): undefined reference to `boost::system::get_generic_category()' a.cpp:(.text+0x668): undefined reference to `boost::system::get_generic_category()' a.cpp:(.text+0x672): undefined reference to `boost::system::get_generic_category()' a.cpp:(.text+0x67c): undefined reference to `boost::system::get_system_category()' /tmp/ccCv9ZJA.o: In function `boost::system::error_code::error_code()': a.cpp:(.text._ZN5boost6system10error_codeC2Ev[_ZN5boost6system10error_codeC5Ev]+0x10): undefined reference to `boost::system::get_system_category()' /tmp/ccCv9ZJA.o: In function `boost::asio::error::get_system_category()': a.cpp:(.text._ZN5boost4asio5error19get_system_categoryEv[boost::asio::error::get_system_category()]+0x7): undefined reference to `boost::system::get_system_category()' /tmp/ccCv9ZJA.o: In function `boost::asio::detail::posix_thread::~posix_thread()': a.cpp:(.text._ZN5boost4asio6detail12posix_threadD2Ev[_ZN5boost4asio6detail12posix_threadD5Ev]+0x1d): undefined reference to `pthread_detach' /tmp/ccCv9ZJA.o: In function `boost::asio::detail::posix_thread::join()': a.cpp:(.text._ZN5boost4asio6detail12posix_thread4joinEv[boost::asio::detail::posix_thread::join()]+0x25): undefined reference to `pthread_join' /tmp/ccCv9ZJA.o: In function `boost::asio::detail::posix_tss_ptr<boost::asio::detail::call_stack<boost::asio::detail::task_io_service<boost::asio::detail::epoll_reactor<false> > >::context>::~posix_tss_ptr()': a.cpp:(.text._ZN5boost4asio6detail13posix_tss_ptrINS1_10call_stackINS1_15task_io_serviceINS1_13epoll_reactorILb0EEEEEE7contextEED2Ev[_ZN5boost4asio6detail13posix_tss_ptrINS1_10call_stackINS1_15task_io_serviceINS1_13epoll_reactorILb0EEEEEE7contextEED5Ev]+0xf): undefined reference to `pthread_key_delete' /tmp/ccCv9ZJA.o: In function `boost::asio::detail::posix_tss_ptr<boost::asio::detail::call_stack<boost::asio::detail::task_io_service<boost::asio::detail::epoll_reactor<false> > >::context>::posix_tss_ptr()': a.cpp:(.text._ZN5boost4asio6detail13posix_tss_ptrINS1_10call_stackINS1_15task_io_serviceINS1_13epoll_reactorILb0EEEEEE7contextEEC2Ev[_ZN5boost4asio6detail13posix_tss_ptrINS1_10call_stackINS1_15task_io_serviceINS1_13epoll_reactorILb0EEEEEE7contextEEC5Ev]+0x22): undefined reference to `pthread_key_create' collect2: ld returned 1 exit status How can I fix it? Thank you.

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• how to display these text on blackberry and how to show the hyperlinks

  - by Changqi

  <!DOCTYPE html PUBLIC "-//W3C//DTD HTML 4.01 Strict//EN" "http://www.w3.org/TR/html4/strict.dtd"> <html> <head> <meta http-equiv="Content-Type" content="text/html; charset=UTF-8"> <title>A history of Canoe Cove /</title> </head> <body> <div class="tei"> <p> A History of </p> <p> The General Stores </p> <p> There were several general stores in our <a class="search orgName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.orgNameTERM:%22Cove%22+AND+dc.type:collection">Cove</a> at different times. The one that lasted longest was at the Corner across from the school and it had many owners. Who established it is unclear but John MacKenzie, the piper, who was also a shoe maker lived there. He was a relative of the present day MacKenzies of <a class="search placeName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.placeNameTERM:%22Canoe Cove%22+AND+dc.type:collection">Canoe Cove</a>. William MacKay who married Christena MacLean was operating it when it burned down and a store which had belonged to Neil "Cooper" MacLean was moved across to the site. This was later bought by <span class="persName"><a class="search persName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.persNameTERM:%22MacCannell+Neil%22+AND+dc.type:collection"> Neil MacCannell </a></span> of Long Creek , a schoolteacher who taught in the <a class="search orgName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.orgNameTERM:%22Cove%22+AND+dc.type:collection">Cove</a> for a few years. <span class="persName"><a class="search persName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.persNameTERM:%22MacNevin+Hector%22+AND+dc.type:collection"> Hector MacNevin </a></span> from <a class="search placeName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.placeNameTERM:%22St. Catherines%22+AND+dc.type:collection">St. Catherines</a> operated it for a year while it still belonged to <span class="persName"><a class="search persName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.persNameTERM:%22MacCannell+Neil%22+AND+dc.type:collection"> Neil MacCannell </a></span> because Neil had accepted a job in <a class="search placeName" target="_blank" href="http://islandlives.net/fedora/ilives_book_search/tei.placeNameTERM:%22Charlottetown%22+AND+dc.type:collection">Charlottetown</a> as clerk of the Court. Later Mrs. John Angus Darrach bought it and she and her son George ran it for years until both had health problems, and had to close the store after which closing it never reopened. After George died and his wife Hazel moved to Montague to live with her family the building was sold to Robert Patterson . Rob lived in it for a few years, making many improvements then sold it to Kirk McAleer. </p> </div>

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• Does anyone know how to appropriately deal with user timezones in rails 2.3?

  - by Amazing Jay

  We're building a rails app that needs to display dates (and more importantly, calculate them) in multiple timezones. Can anyone point me towards how to work with user timezones in rails 2.3(.5 or .8) The most inclusive article I've seen detailing how user time zones are supposed to work is here: http://wiki.rubyonrails.org/howtos/time-zones... although it is unclear when this was written or for what version of rails. Specifically it states that: "Time.zone - The time zone that is actually used for display purposes. This may be set manually to override config.time_zone on a per-request basis." Keys terms being "display purposes" and "per-request basis". Locally on my machine, this is true. However on production, neither are true. Setting Time.zone persists past the end of the request (to all subsequent requests) and also affects the way AR saves to the DB (basically treating any date as if it were already in UTC even when its not), thus saving completely inappropriate values. We run Ruby Enterprise Edition on production with passenger. If this is my problem, do we need to switch to JRuby or something else? To illustrate the problem I put the following actions in my ApplicationController right now: def test p_time = Time.now.utc s_time = Time.utc(p_time.year, p_time.month, p_time.day, p_time.hour) logger.error "TIME.ZONE" + Time.zone.inspect logger.error ENV['TZ'].inspect logger.error p_time.inspect logger.error s_time.inspect jl = JunkLead.create! jl.date_at = s_time logger.error s_time.inspect logger.error jl.date_at.inspect jl.save! logger.error s_time.inspect logger.error jl.date_at.inspect render :nothing => true, :status => 200 end def test2 Time.zone = 'Mountain Time (US & Canada)' logger.error "TIME.ZONE" + Time.zone.inspect logger.error ENV['TZ'].inspect render :nothing => true, :status => 200 end def test3 Time.zone = 'UTC' logger.error "TIME.ZONE" + Time.zone.inspect logger.error ENV['TZ'].inspect render :nothing => true, :status => 200 end and they yield the following: Processing ApplicationController#test (for 98.202.196.203 at 2010-12-24 22:15:50) [GET] TIME.ZONE#<ActiveSupport::TimeZone:0x2c57a68 @tzinfo=#<TZInfo::DataTimezone: Etc/UTC>, @name="UTC", @utc_offset=0> nil Fri Dec 24 22:15:50 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 22:00:00 UTC +00:00 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 22:00:00 UTC +00:00 Completed in 21ms (View: 0, DB: 4) | 200 OK [http://www.dealsthatmatter.com/test] Processing ApplicationController#test2 (for 98.202.196.203 at 2010-12-24 22:15:53) [GET] TIME.ZONE#<ActiveSupport::TimeZone:0x2c580a8 @tzinfo=#<TZInfo::DataTimezone: America/Denver>, @name="Mountain Time (US & Canada)", @utc_offset=-25200> nil Completed in 143ms (View: 1, DB: 3) | 200 OK [http://www.dealsthatmatter.com/test2] Processing ApplicationController#test (for 98.202.196.203 at 2010-12-24 22:15:59) [GET] TIME.ZONE#<ActiveSupport::TimeZone:0x2c580a8 @tzinfo=#<TZInfo::DataTimezone: America/Denver>, @name="Mountain Time (US & Canada)", @utc_offset=-25200> nil Fri Dec 24 22:15:59 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 15:00:00 MST -07:00 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 15:00:00 MST -07:00 Completed in 20ms (View: 0, DB: 4) | 200 OK [http://www.dealsthatmatter.com/test] Processing ApplicationController#test3 (for 98.202.196.203 at 2010-12-24 22:16:03) [GET] TIME.ZONE#<ActiveSupport::TimeZone:0x2c57a68 @tzinfo=#<TZInfo::DataTimezone: Etc/UTC>, @name="UTC", @utc_offset=0> nil Completed in 17ms (View: 0, DB: 2) | 200 OK [http://www.dealsthatmatter.com/test3] Processing ApplicationController#test (for 98.202.196.203 at 2010-12-24 22:16:04) [GET] TIME.ZONE#<ActiveSupport::TimeZone:0x2c57a68 @tzinfo=#<TZInfo::DataTimezone: Etc/UTC>, @name="UTC", @utc_offset=0> nil Fri Dec 24 22:16:05 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 22:00:00 UTC +00:00 Fri Dec 24 22:00:00 UTC 2010 Fri, 24 Dec 2010 22:00:00 UTC +00:00 Completed in 151ms (View: 0, DB: 4) | 200 OK [http://www.dealsthatmatter.com/test] It should be clear above that the 2nd call to /test shows Time.zone set to Mountain, even though it shouldn't. Additionally, checking the database reveals that the test action when run after test2 saved a JunkLead record with a date of 2010-12-22 15:00:00, which is clearly wrong.

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• jquery mouse events masked by img in IE8 - all other platforms/browsers work fine

  - by Bruce

  Using the jQuery mouse events for mouseenter, mouseleave or hover all work swimmingly on all Mac or Windows browsers except for IE8/Windows. I am using mouseenter and mouseleave to test for hot rectangles (absolutely positioned and dimensioned divs that have no content) over an image used as hotspots to make visible the navigation buttons when different regions of the main enclosing rectangle (image) are touched by the cursor. Windows/IE jQuery never sends notifications (mouseenter our mouseleave) as the cursor enters or exits one of the target divs. If I turn off the visibility of the image everything works as expected (like it does in every other browser), so the image is effectively blocking all messages (the intention was for the image to be a background and all the divs float above it, where they can be clicked on). I understand that there's a z-index gotcha (so explicitly specifying z-index for each absolute positioned div does not work), but unclear as to how to nest or order multiple divs to allow a single set of jQuery rules to support all browsers. The image tag seems to trump all other divs and always appear in front of them... BTW: I could not use i m g as a tag in this text so it is spelled image in the following, so the input editor does not think that I am trying to pull a fast one on stackoverflow... How used? "mainview" is the background image, "zoneleft" and "zoneright" are the active areas where when the cursor enters the nav buttons "leftarrow" and rightarrow" are supposed to appear. Javascript $("div#zoneleft").bind("mouseenter",function () // enters left zone see left arrow { arrowVisibility("left"); document.getElementById("leftarrow").style.display="block"; }).bind("mouseleave",function () { document.getElementById("leftarrow").style.visibility="hidden"; document.getElementById("rightarrow").style.visibility="hidden"; }); HTML <div id="zoneleft" style="position:absolute; top:44px; left:0px; width:355px; height:372px; z-index:40;"> <div id="leftarrow" style="position:absolute; top:158px; left:0px; z-index:50;"><img src="images/newleft.png" width="59" height="56"/></div></div> <div id="zoneright" style="position:absolute; top:44px; left:355px; width:355px; height:372px; z-index:40;"> <div id="rightarrow" style="position:absolute; top:158px; left:296px; z-index:50;"> (tag named changed so that I could include it here) <image src="images/newright.png" width="59" height="56" /></div></div> </div><!-- navbuttons --> <image id="mainview" style="z-index:-1;" src="images/projectPhotos/photo1.jpg" width:710px; height:372px; /> (tag named changed so that I could include it here) </div><!--photo-->

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• Associating an Object with other Objects and Properties of those Objects

  - by alzoid

  I am looking for some help with designing some functionality in my application. I already have something similar designed but this problem is a little different. Background: In my application we have different Modules. Data in each module can be associated to other modules. Each Module is represented by an Object in our application. Module 1 can be associated with Module 2 and Module 3. Currently I use a factory to provide the proper DAO for getting and saving this data. It looks something like this: class Module1Factory { public static Module1BridgeDAO createModule1BridgeDAO(int moduleid) { switch (moduleId) { case Module.Module2Id: return new Module1_Module2DAO(); case Module.Module3Id: return new Module1_Module3DAO(); default: return null; } } } Module1_Module2 and Module1_Module3 implement the same BridgeModule interface. In the database I have a Table for every module (Module1, Module2, Module3). I also have a bridge table for each module (they are many to many) Module1_Module2, Module1_Module3 etc. The DAO basically handles all code needed to manage the association and retrieve its own instance data for the calling module. Now when we add new modules that associate with Module1 we simply implement the ModuleBridge interface and provide the common functionality. New Development We are adding a new module that will have the ability to be associated with other Modules as well as specific properties of that module. The module is basically providing the user the ability to add their custom forms to our other modules. That way they can collect additional information along with what we provide. I want to start associating my Form module with other modules and their properties. Ie if Module1 has a property Category, I want to associate an instance From data with that property. There are many Forms. If a users creates an instance of Module2, they may always want to also have certain form(s) attached to that Module2 instance. If they create an instance of Module2 and select Category 1, then I may want additional Form(s) created. I prototyped something like this: Form FormLayout (contains the labels and gui controls) FormModule (associates a form with all instances of a module) Form Instance (create an instance of a form to be filled out) As I thought about it I was thinking about making a new FormModule table/class/dao for each Module and Property that I add. So I might have: FormModule1 FormModule1Property1 FormModule1Property2 FormModule1Property3 FormModule1Property4 FormModule2 FormModule3 FormModule3Property1 Then as I did previously, I would use a factory to get the proper DAO for dealing with all of these. I would hand it an array of ids representing different modules and properties and it would return all of the DAOs that I need to call getForms(). Which in turn would return all of the forms for that particular bridge. Some points This will be for a new module so I dont need to expand on the factory code I provided. I just wanted to show an example of what I have done in the past. The new module can be associated with: Other Modules (ie globally for any instance of that module data), Other module properties (ie only if the Module instance has a certian value in one of its properties) I want to make it easy for developers to add associations with other modules and properties easily Can any one suggest any design patterns or strategy's for achieving this? If anything is unclear please let me know. Thank you, Al

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• How to reduce iOS AVPlayer start delay

  - by Bernt Habermeier

  Note, for the below question: All assets are local on the device -- no network streaming is taking place. The videos contain audio tracks. I'm working on an iOS application that requires playing video files with minimum delay to start the video clip in question. Unfortunately we do not know what specific video clip is next until we actually need to start it up. Specifically: When one video clip is playing, we will know what the next set of (roughly) 10 video clips are, but we don't know which one exactly, until it comes time to 'immediately' play the next clip. What I've done to look at actual start delays is to call addBoundaryTimeObserverForTimes on the video player, with a time period of one millisecond to see when the video actually started to play, and I take the difference of that time stamp with the first place in the code that indicates which asset to start playing. From what I've seen thus-far, I have found that using the combination of AVAsset loading, and then creating an AVPlayerItem from that once it's ready, and then waiting for AVPlayerStatusReadyToPlay before I call play, tends to take between 1 and 3 seconds to start the clip. I've since switched to what I think is roughly equivalent: calling [AVPlayerItem playerItemWithURL:] and waiting for AVPlayerItemStatusReadyToPlay to play. Roughly same performance. One thing I'm observing is that the first AVPlayer item load is slower than the rest. Seems one idea is to pre-flight the AVPlayer with a short / empty asset before trying to play the first video might be of good general practice. [http://stackoverflow.com/questions/900461/slow-start-for-avaudioplayer-the-first-time-a-sound-is-played] I'd love to get the video start times down as much as possible, and have some ideas of things to experiment with, but would like some guidance from anyone that might be able to help. Update: idea 7, below, as-implemented yields switching times of around 500 ms. This is an improvement, but it it'd be nice to get this even faster. Idea 1: Use N AVPlayers (won't work) Using ~ 10 AVPPlayer objects and start-and-pause all ~ 10 clips, and once we know which one we really need, switch to, and un-pause the correct AVPlayer, and start all over again for the next cycle. I don't think this works, because I've read there is roughly a limit of 4 active AVPlayer's in iOS. There was someone asking about this on StackOverflow here, and found out about the 4 AVPlayer limit: fast-switching-between-videos-using-avfoundation Idea 2: Use AVQueuePlayer (won't work) I don't believe that shoving 10 AVPlayerItems into an AVQueuePlayer would pre-load them all for seamless start. AVQueuePlayer is a queue, and I think it really only makes the next video in the queue ready for immediate playback. I don't know which one out of ~10 videos we do want to play back, until it's time to start that one. ios-avplayer-video-preloading Idea 3: Load, Play, and retain AVPlayerItems in background (not 100% sure yet -- but not looking good) I'm looking at if there is any benefit to load and play the first second of each video clip in the background (suppress video and audio output), and keep a reference to each AVPlayerItem, and when we know which item needs to be played for real, swap that one in, and swap the background AVPlayer with the active one. Rinse and Repeat. The theory would be that recently played AVPlayer/AVPlayerItem's may still hold some prepared resources which would make subsequent playback faster. So far, I have not seen benefits from this, but I might not have the AVPlayerLayer setup correctly for the background. I doubt this will really improve things from what I've seen. Idea 4: Use a different file format -- maybe one that is faster to load? I'm currently using .m4v's (video-MPEG4) H.264 format. I have not played around with other formats, but it may well be that some formats are faster to decode / get ready than others. Possible still using video-MPEG4 but with a different codec, or maybe quicktime? Maybe a lossless video format where decoding / setup is faster? Idea 5: Combination of lossless video format + AVQueuePlayer If there is a video format that is fast to load, but maybe where the file size is insane, one idea might be to pre-prepare the first 10 seconds of each video clip with a version that is boated but faster to load, but back that up with an asset that is encoded in H.264. Use an AVQueuePlayer, and add the first 10 seconds in the uncompressed file format, and follow that up with one that is in H.264 which gets up to 10 seconds of prepare/preload time. So I'd get 'the best' of both worlds: fast start times, but also benefits from a more compact format. Idea 6: Use a non-standard AVPlayer / write my own / use someone else's Given my needs, maybe I can't use AVPlayer, but have to resort to AVAssetReader, and decode the first few seconds (possibly write raw file to disk), and when it comes to playback, make use of the raw format to play it back fast. Seems like a huge project to me, and if I go about it in a naive way, it's unclear / unlikely to even work better. Each decoded and uncompressed video frame is 2.25 MB. Naively speaking -- if we go with ~ 30 fps for the video, I'd end up with ~60 MB/s read-from-disk requirement, which is probably impossible / pushing it. Obviously we'd have to do some level of image compression (perhaps native openGL/es compression formats via PVRTC)... but that's kind crazy. Maybe there is a library out there that I can use? Idea 7: Combine everything into a single movie asset, and seekToTime One idea that might be easier than some of the above, is to combine everything into a single movie, and use seekToTime. The thing is that we'd be jumping all around the place. Essentially random access into the movie. I think this may actually work out okay: avplayer-movie-playing-lag-in-ios5 Which approach do you think would be best? So far, I've not made that much progress in terms of reducing the lag.

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