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  • How do you select form elements in JQuery based upon an html table?

    - by Swoop
    I am working on some ASP.NET web forms which involves some dynamic generation, and I need to add some onClick helpers on the client side. I have a basic outline of something working, except for one huge problem. There are multiple HTML tables, each generated by a different ASP.NET web control. Each table can contain overlapping field names, which is causing a problem with my JQuery click event handlers. The click event handler is linking to unintended form fields in addition to the intended form field. I have provided a simplified sample version of the code below. This code is trying to set the value of textbox box1 when a particular radiobutton is selected in the table with id=thing1. Obviously, the jquery code will be triggered for the form fields in both tables. The tables are dynamically added to the webpage based upon different conditions. It is possible that no tables will be loaded, only 1 table, or both tables might load. In the future, other tables could be added. Each table comes from a different .net web control. Other than renaming the form fields to make sure they are unique across all user controls, is there a way to have JQuery act only on the intended form fields? In other words, could the table ID be incorporated into the JQuery code in a manner that does not become a nightmare to maintain later? <script> $(document).ready(function() { $("[id$=radio1_0]").click(function() { $("[id$=box1]").attr("value", ""); }); $("[id$=radio1_1]").click(function() { $("[id$=box1]").attr("value", "N/A"); }); </script> <table id="thing1"> <tr><td> <radiobuttonlist id="radio1"/> <listitem>yes</listitem> <listitem>no</listitem> </td></tr> <tr><td> <textbox id="box1"/> </td></tr> </table> <table id="thing2"> <tr><td> <radiobuttonlist id="radio1"/> <listitem>yes</listitem> <listitem>no</listitem> </td></tr> <tr><td> <textbox id="box1"/> </tr></td> </table>

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  • Robust way to save/load objects with dependencies?

    - by mrteacup
    I'm writing an Android game in Java and I need a robust way to save and load application state quickly. The question seems to apply to most OO languages. To understand what I need to save: I'm using a Strategy pattern to control my game entities. The idea is I have a very general Entity class which e.g. stores the location of a bullet/player/enemy and I then attach a Behaviour class that tells the entity how to act: class Entiy { float x; float y; Behavior b; } abstract class Behavior { void update(Entity e); {} // Move about at a constant speed class MoveBehavior extends Behavior { float speed; void update ... } // Chase after another entity class ChaseBehavior extends Behavior { Entity target; void update ... } // Perform two behaviours in sequence class CombineBehavior extends Behavior { Behaviour a, b; void update ... } Essentially, Entity objects are easy to save but Behaviour objects can have a semi-complex graph of dependencies between other Entity objects and other Behaviour objects. I also have cases where a Behaviour object is shared between entities. I'm willing to change my design to make saving/loading state easier, but the above design works really well for structuring the game. Anyway, the options I've considered are: Use Java serialization. This is meant to be really slow in Android (I'll profile it sometime). I'm worried about robustness when changes are made between versions however. Use something like JSON or XML. I'm not sure how I would cope with storing the dependencies between objects however. Would I have to give each object a unique ID and then use these IDs on loading to link the right objects together? I thought I could e.g. change the ChaseBehaviour to store a ID to an entity, instead of a reference, that would be used to look up the Entity before performing the behaviour. I'd rather avoid having to write lots of loading/saving code myself as I find it really easy to make mistakes (e.g. forgetting to save something, reading things out in the wrong order). Can anyone give me any tips on good formats to save to or class designs that make saving state easier?

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  • string s; &s+1; Legal? UB?

    - by John Dibling
    Consider the following code: #include <cstdlib> #include <iostream> #include <string> #include <vector> #include <algorithm> using namespace std; int main() { string myAry[] = { "Mary", "had", "a", "Little", "Lamb" }; const size_t numStrs = sizeof(myStr)/sizeof(myAry[0]); vector<string> myVec(&myAry[0], &myAry[numStrs]); copy( myVec.begin(), myVec.end(), ostream_iterator<string>(cout, " ")); return 0; } Of interest here is &myAry[numStrs]: numStrs is equal to 5, so &myAry[numStrs] points to something that doesn't exist; the sixth element in the array. There is another example of this in the above code: myVec.end(), which points to one-past-the-end of the vector myVec. It's perfecly legal to take the address of this element that doesn't exist. We know the size of string, so we know where the address of the 6th element of a C-style array of strings must point to. So long as we only evaluate this pointer and never dereference it, we're fine. We can even compare it to other pointers for equality. The STL does this all the time in algorithms that act on a range of iterators. The end() iterator points past the end, and the loops keep looping while a counter != end(). So now consider this: #include <cstdlib> #include <iostream> #include <string> #include <vector> #include <algorithm> using namespace std; int main() { string myStr = "Mary"; string* myPtr = &myStr; vector<string> myVec2(myPtr, &myPtr[1]); copy( myVec2.begin(), myVec2.end(), ostream_iterator<string>(cout, " ")); return 0; } Is this code legal and well-defined? It is legal and well-defined to take the address of an array element past the end, as in &myAry[numStrs], so should it be legal and well-defined to pretend that myPtr is also an array?

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  • how can we change the value by using radio buttons

    - by magna
    I am making a website in Adobe Dreamweaver with php. In the site there’s a 3 buttons for selecting payment method that will act as the continue button. What I want is when the user checks a radio buttons (I agree button), it will be add with that amount and display with previous amount.. there is three buttons which has the corresponding values(amount in pounds).. plz check my website http://www.spsmobile.co.uk in this linkgo to mobile phone unlocking and after add the cart click make payment it will go to next page there is a delivery mail details.. for that delivery mail details only am asking.. here i mentioned code: <input id="radio-1" type="radio" name="rmr" value="1"> <label for="radio-1">£3</label> <input id="radio-2" type="radio" name="rmr" value="2"> <label for="radio-2">£5.5</label> <input id="radio-3" type="radio" name="rmr" value="4"> <label for="radio-3">£10</label> <div class="total-text" style="font-size:36px">£10</div> var total = parseInt($("div.total-text").text().substring(1), 10); $("input[name='rmr']").bind('change', function() { var amount = 0; switch (this.value) { case "1": amount = 3; break; case "2": amount = 5.5; break; case "4": amount = 10; break; } $("div.total-text").text("£" + (total + amount)); }); but there is no change , my previous amount did not add with that. while am clicking previous amount only displayed on browser.. i need when i cliks radio button the value should change correspondingly.. where i did that mistake...plz give me some idea and what should i do..is there any need for storing db.. thanks in adv

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  • How to reflect over T to build an expression tree for a query?

    - by Alex
    Hi all, I'm trying to build a generic class to work with entities from EF. This class talks to repositories, but it's this class that creates the expressions sent to the repositories. Anyway, I'm just trying to implement one virtual method that will act as a base for common querying. Specifically, it will accept a an int and it only needs to perform a query over the primary key of the entity in question. I've been screwing around with it and I've built a reflection which may or may not work. I say that because I get a NotSupportedException with a message of LINQ to Entities does not recognize the method 'System.Object GetValue(System.Object, System.Object[])' method, and this method cannot be translated into a store expression. So then I tried another approach and it produced the same exception but with the error of The LINQ expression node type 'ArrayIndex' is not supported in LINQ to Entities. I know it's because EF will not parse the expression the way L2S will. Anyway, I'm hopping someone with a bit more experience can point me into the right direction on this. I'm posting the entire class with both attempts I've made. public class Provider<T> where T : class { protected readonly Repository<T> Repository = null; private readonly string TEntityName = typeof(T).Name; [Inject] public Provider( Repository<T> Repository) { this.Repository = Repository; } public virtual void Add( T TEntity) { this.Repository.Insert(TEntity); } public virtual T Get( int PrimaryKey) { // The LINQ expression node type 'ArrayIndex' is not supported in // LINQ to Entities. return this.Repository.Select( t => (((int)(t as EntityObject).EntityKey.EntityKeyValues[0].Value) == PrimaryKey)).Single(); // LINQ to Entities does not recognize the method // 'System.Object GetValue(System.Object, System.Object[])' method, // and this method cannot be translated into a store expression. return this.Repository.Select( t => (((int)t.GetType().GetProperties().Single( p => (p.Name == (this.TEntityName + "Id"))).GetValue(t, null)) == PrimaryKey)).Single(); } public virtual IList<T> GetAll() { return this.Repository.Select().ToList(); } protected virtual void Save() { this.Repository.Update(); } }

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  • form inside tabview doesn't work

    - by user3536737
    i am working with jsf and primefaces , and here is what 've tried well i want to creat a tabview that get data from an arraylist in my bean i get for exemple 4 tabs , and inside each one i've created a hidden panel where i have a form with 2 input text to update informations , do i display the panel when i click on the second button Update , after that my panel is not hidden anymore , and i set the new values and click on the second button to update the informations , the problem is that the updating and the execution is working only for the first tab , it means when i try to update the new informations it works for the first one and for the other tabs it doesn't here is the code <p:tab title="#{rr.nom_ressource}"> <h:panelGrid> <h:graphicImage value="Ressources/images/emp.jpg" style="vertical-align:middle" /> <span style="font-size:15px; width:170px; display:inline-block;"> Nom : #{rr.nom_ressource} Type: #{rr.type_ressource} Specification: #{rr.experience} </span> <h:commandButton image="Ressources/images/delete.jpg" actionListener="#{SelectBean.act}" update=":form" style="vertical-align:middle" > Update </h:commandButton> <h:commandButton update=":outPanel" actionListener="#{SelectBean.mod1()}" image="Ressources/images/update.png" style="vertical-align:middle" > Modifier </h:commandButton> <h:form id="form111"> <p:growl id="growl" showDetail="true" sticky="true" /> <p:panel rendered ="#{SelectBean.bol}" closable="true" toggleable="true" id="outPanel" styleClass="outPanel" widgetVar="outpanel"> <h:outputLabel value="Nom " /> <h:inputText value="#{SelectBean.nom}" /> <br/> <h:outputLabel value="Experience " /> <h:inputText value="#{SelectBean.exp}" /> <br/> <h:commandButton value="Update" action="#{SelectBean.done}"/> </p:panel> </h:form> </h:panelGrid> </p:tab> for my managedbean the code is correct i think the problem is here

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  • 4 table query / join. getting duplicate rows

    - by Horse
    So I have written a query that will grab an order (this is for an ecommerce type site), and from that order id it will get all order items (ecom_order_items), print options (c_print_options) and images (images). The eoi_p_id is currently a foreign key from the images table. This works fine and the query is: SELECT eoi_parentid, eoi_p_id, eoi_po_id, eoi_quantity, i_id, i_parentid, po_name, po_price FROM ecom_order_items, images, c_print_options WHERE eoi_parentid = '1' AND i_id = eoi_p_id AND po_id = eoi_po_id; The above would grab all the stuff I need for order #1 Now to complicate things I added an extra table (ecom_products), which needs to act in a similar way to the images table. The eoi_p_id can also point at a foreign key in this table too. I have added an extra field 'eoi_type' which will either have the value 'image', or 'product'. Now items in the order could be made up of a mix of items from images or ecom_products. Whatever I try it either ends up with too many records, wont actually output any with eoi_type = 'product', and just generally wont work. Any ideas on how to achieve what I am after? Can provide SQL samples if needed? SELECT eoi_id, eoi_parentid, eoi_p_id, eoi_po_id, eoi_po_id_2, eoi_quantity, eoi_type, i_id, i_parentid, po_name, po_price, po_id, ep_id FROM ecom_order_items, images, c_print_options, ecom_products WHERE eoi_parentid = '9' AND i_id = eoi_p_id AND po_id = eoi_po_id The above outputs duplicate rows and doesnt work as expected. Am I going about this the wrong way? Should I have seperate foreign key fields for the eoi_p_id depending it its an image or a product? Should I be using JOINs? Here is a mysql explain of the tables in question ecom_products +-------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +-------------+--------------+------+-----+---------+----------------+ | ep_id | int(8) | NO | PRI | NULL | auto_increment | | ep_title | varchar(255) | NO | | NULL | | | ep_link | text | NO | | NULL | | | ep_desc | text | NO | | NULL | | | ep_imgdrop | text | NO | | NULL | | | ep_price | decimal(6,2) | NO | | NULL | | | ep_category | varchar(255) | NO | | NULL | | | ep_hide | tinyint(1) | NO | | 0 | | | ep_featured | tinyint(1) | NO | | 0 | | +-------------+--------------+------+-----+---------+----------------+ ecom_order_items +--------------+-------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +--------------+-------------+------+-----+---------+----------------+ | eoi_id | int(8) | NO | PRI | NULL | auto_increment | | eoi_parentid | int(8) | NO | | NULL | | | eoi_type | varchar(32) | NO | | NULL | | | eoi_p_id | int(8) | NO | | NULL | | | eoi_po_id | int(8) | NO | | NULL | | | eoi_quantity | int(4) | NO | | NULL | | +--------------+-------------+------+-----+---------+----------------+ c_print_options +------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +------------+--------------+------+-----+---------+----------------+ | po_id | int(8) | NO | PRI | NULL | auto_increment | | po_name | varchar(255) | NO | | NULL | | | po_price | decimal(6,2) | NO | | NULL | | +------------+--------------+------+-----+---------+----------------+ images +--------------+--------------+------+-----+---------+----------------+ | Field | Type | Null | Key | Default | Extra | +--------------+--------------+------+-----+---------+----------------+ | i_id | int(8) | NO | PRI | NULL | auto_increment | | i_filename | varchar(255) | NO | | NULL | | | i_data | longtext | NO | | NULL | | | i_parentid | int(8) | NO | | NULL | | +--------------+--------------+------+-----+---------+----------------+

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  • Oracle performance problem

    - by jreid42
    We are using an Oracle 11G machine that is very powerful; has redundant storage etc. It's a beast from what I have been told. We just got this DB for a tool that when I first came on as a coop had like 20 people using, now its upwards of 150 people. I am the only one working on it :( We currently have a system in place that distributes PERL scripts across our entire data center essentially giving us a sort of "grid" computing power. The Perl scripts run a sort of simulation and report back the results to the database. They do selects / inserts. The load is not very high for each script but it could be happening across 20-50 systems at the same time. We then have multiple data centers and users all hitting the same database with this same approach. Our main problem with this is that our database is getting overloaded with connections and having to drop some. We sometimes have upwards of 500 connections. These are old perl scripts and they do not handle this well. Essentially they fail and the results are lost. I would rather avoid having to rewrite a lot of these as they are poorly written, and are a headache to even look at. The database itself is not overloaded, just the connection overhead is too high. We open a connection, make a quick query and then drop the connection. Very short connections but many of them. The database team has basically said we need to lower the number of connections or they are going to ignore us. Because this is distributed across our farm we cant implement persistent connections. I do this with our webserver; but its on a fixed system. The other ones are perl scripts that get opened and closed by the distribution tool and thus arent always running. What would be my best approach to resolving this issue? The scripts themselves can wait for a connection to be open. They do not need to act immediately. Some sort of queing system? I've been suggested to set up a few instances of a tool called "SQL Relay". Maybe one in each data center. How reliable is this tool? How good is this approach? Would it work for what we need? We could have one for each data center and relay requests through it to our main database, keeping a pipeline of open persistent connections? Does this make sense? Is there any other suggestions you can make? Any ideas? Any help would be greatly appreciated. Sadly I am just a coop student working for a very big company and somehow all of this has landed all on my shoulders (there is literally nobody to ask for help; its a hardware company, everybody is hardware engineers, and the database team is useless and in India) and I am quite lost as what the best approach would be? I am extremely overworked and this problem is interfering with on going progress and basically needs to be resolved as quickly as possible; preferably without rewriting the whole system, purchasing hardware (not gonna happen), or shooting myself in the foot. HELP LOL!

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  • How do you handle authentication across domains?

    - by William Ratcliff
    I'm trying to save users of our services from having to have multiple accounts/passwords. I'm in a large organization and there's one group that handles part of user authentication for users who are from outside the facility (primarily for administrative functions). They store a secure cookie to establish a session and communicate only via HTTPS via the browser. Sessions expire either through: 1) explicit logout of the user 2) Inactivity 3) Browser closes My team is trying to write a web application to help users analyze data that they've taken (or are currently taking) while at our facility. We need to determine if a user is 1) authenticated 2) Some identifier for that user so we can store state for them (what analysis they are working on, etc.) So, the problem is how do you authenticate across domains (the authentication server for the other application lives in a border region between public and private--we will live in the public region). We have come up with some scenarios and I'd like advice about what is best practice, or if there is one we haven't considered. Let's start with the case where the user is authenticated with the authentication server. 1) The authentication server leaves a public cookie in the browser with their primary key for a user. If this is deemed sensitive, they encrypt it on their server and we have the key to decrypt it on our server. When the user visits our site, we check for this public cookie. We extract the user_id and use a public api for the authentication server to request if the user is logged in. If they are, they send us a response with: response={ userid :we can then map this to our own user ids. If necessary, we can request additional information such as email-address/display name once (to notify them if long running jobs are done, or to share results with other people, like with google_docs). account_is_active:Make sure that the account is still valid session_is_active: Is their session still active? If we query this for a valid user, this will have a side effect that we will reset the last_time_session_activated value and thus prolong their session with the authentication server last_time_session_activated: let us know how much time they have left ip_address_session_started_from:make sure the person at our site is coming from the same ip as they started the session at } Given this response, we either accept them as authenticated and move on with our app, or redirect them to the login page for the authentication server (question: if we give an encrypted portion of the response (signed by us) with the page to redirect them to, do we open any gaping security holes in the authentication server)? The flaw that we've found with this is that if the user visits evilsite.com and they look at the session cookie and send a query to the public api of the authentication server, they can keep the session alive and if our original user leaves the machine without logging out, then the next user will be able to access their session (this was possible before, but having the session alive eternally makes this worse). 2) The authentication server redirects all requests made to our domain to us and we send responses back through them to the user. Essentially, they act as a proxy. The advantage of this is that we can handshake with the authentication server, so it's safe to be trusted with the email address/name of the user and they don't have to reenter it So, if the user tries to go to: authentication_site/mysite_page1 they are redirected to mysite. Which would you choose, or is there a better way? The goal is to minimize the "Yet Another Password/Yet another username" problem... Thanks!!!!

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  • Increase samba space on open suse 12.1

    - by Kapil Sharma
    I know linux basics but not an expert. IT guy left the job here and there is some time before new hire. So sorry if question is very basic. We have local testing server based on Open SUSE 12.1, which also act as shared drive between dev/mgmt team here and using Samba for that. Now we are running out of space on samba, even though server's 2*1TB harddisk is nearly 90% free. My question is, what is limiting Samba and how can I increase its limit? We need around at least 500 GB as shared drive but currently its just 25 GB. I don't need step by step answer, just a link to any helpful article would be sufficient. Probably I'm putting wrong keywords in google so not getting any helpful link. EDIT: Output of commands in the first comment. All commands were run as root user df -h (getting error with df -ht) Filesystem Size Used Avail Use% Mounted on rootfs 30G 5.1G 23G 19% / devtmpfs 2.0G 36K 2.0G 1% /dev tmpfs 2.0G 1.1M 2.0G 1% /dev/shm tmpfs 2.0G 676K 2.0G 1% /run /dev/sda2 30G 5.1G 23G 19% / tmpfs 2.0G 0 2.0G 0% /sys/fs/cgroup tmpfs 2.0G 676K 2.0G 1% /var/run tmpfs 2.0G 0 2.0G 0% /media tmpfs 2.0G 676K 2.0G 1% /var/lock /dev/sda3 36G 31G 3.3G 91% /home fdisk -l /dev/[hmsv]d* Disk /dev/sda: 80.0 GB, 80026361856 bytes 255 heads, 63 sectors/track, 9729 cylinders, total 156301488 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x2d4a2d49 Device Boot Start End Blocks Id System /dev/sda1 2048 16771071 8384512 82 Linux swap / Solaris /dev/sda2 * 16771072 79681535 31455232 83 Linux /dev/sda3 79681536 156301311 38309888 83 Linux Disk /dev/sda1: 8585 MB, 8585740288 bytes 255 heads, 63 sectors/track, 1043 cylinders, total 16769024 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00000000 Disk /dev/sda1 doesn't contain a valid partition table Disk /dev/sda2: 32.2 GB, 32210157568 bytes 255 heads, 63 sectors/track, 3915 cylinders, total 62910464 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00000000 Device Boot Start End Blocks Id System Disk /dev/sda3: 39.2 GB, 39229325312 bytes 255 heads, 63 sectors/track, 4769 cylinders, total 76619776 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x00000000 Disk /dev/sda3 doesn't contain a valid partition table vgs No volume groups found lvs No volume groups found output of vi /etc/samba/smb.conf # smb.conf is the main Samba configuration file. You find a full commented # version at /usr/share/doc/packages/samba/examples/smb.conf.SUSE if the # samba-doc package is installed. # Date: 2011-11-02 [global] workgroup = WORKGROUP passdb backend = tdbsam printing = cups printcap name = cups printcap cache time = 750 cups options = raw map to guest = Bad User include = /etc/samba/dhcp.conf logon path = \\%L\profiles\.msprofile logon home = \\%L\%U\.9xprofile logon drive = P: usershare allow guests = Yes [homes] comment = Home Directories valid users = %S, %D%w%S browseable = No read only = No inherit acls = Yes [profiles] comment = Network Profiles Service path = %H read only = No store dos attributes = Yes create mask = 0600 directory mask = 0700 [users] comment = All users path = /home read only = No inherit acls = Yes veto files = /aquota.user/groups/shares/ [groups] comment = All groups path = /home/groups read only = No inherit acls = Yes [printers] comment = All Printers path = /var/tmp printable = Yes create mask = 0600 browseable = No [print$] comment = Printer Drivers path = /var/lib/samba/drivers write list = @ntadmin root force group = ntadmin create mask = 0664 directory mask = 0775 [allusers] comment = All Users path = /home/shares/allusers valid users = @users force group = users create mask = 0660 directory mask = 0771 writable = yes

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  • Apache cyclic redirection problem

    - by slicedlime
    I have an extremely weird problem with one of my sites. I run a number of blogs off a single apache2 server with a shared wordpress install. Each site has a www.domain.com main domain, but a ServerAlias of domain.com. This works fine for all the blogs except one, which instead of redirecting to www.domain.com redirects to domain.com, causing a cyclic redirection. The configuration for each host looks like this: <VirtualHost *:80> ServerName www.domain.com ServerAlias domain.com DocumentRoot "/home/www/www.domain.com" <Directory "/home/www/www.domain.com"> Options MultiViews Indexes Includes FollowSymLinks ExecCGI AllowOverride All Order allow,deny Allow from all </Directory> </VirtualHost> As this didn't work, I tried a mod_rewrite rule for it, which still didn't redirect correctly. The weird thing here is that if i rewrite it to redirect to any other domain it will redirect correctly, even to another subdomain. So a rewrite rule for domain.com that redirects to foo.domain.com works, but not to www.domain.com. In the same way, trying to redirec to www.domain.com/foo/ ends me up with a redirection to domain.com/foo/. Even weirder, I tried setting up domain.com as a completely separate virtual host, and ran this php test script as index.php on it: <?php header('Location: http://www.domain.com/' . $_SERVER["request_uri"]); ?> Hitting domain.com still redirects to domain.com! Checking out the headers sent to the server verifies that I get exactly the redirect URL I wanted, except with the "www." stripped. The same script works like a charm if I replace www. with foo or redirect to any other domain. This has now foiled me for a long time. I've diffed the vhosts configs for a working domain and the faulty one, and the only difference is the domain name itself. I've diffed the .htaccess files for both sites, and the only difference is a path related to the sharing of wordpress installation for the blogs: php_value include_path ".:/home/www/www.domain.com/local/:/home/www/www.domain.com/" I searched through everything in /etc (including apache conf) for the domain name of the faulty host and found nothing weird, searched through everything in /etc for one of the working ones to make sure it didn't differ, I even went so far to check on the DNS setup of two domains to make sure there wasn't anything strange going on. Here's the response for the faulty one: user@localhost dir $ wget -S domain.com --2010-03-20 21:47:24-- http://domain.com/ Resolving domain.com... x.x.x.x Connecting to domain.com|x.x.x.x|:80... connected. HTTP request sent, awaiting response... HTTP/1.1 301 Moved Permanently Via: 1.1 ISA Connection: Keep-Alive Proxy-Connection: Keep-Alive Content-Length: 0 Date: Sat, 20 Mar 2010 20:47:24 GMT Location: http://domain.com/ Content-Type: text/html; charset=UTF-8 Server: Apache X-Powered-By: PHP/5.2.10-pl0-gentoo X-Pingback: http://domain.com/xmlrpc.php Keep-Alive: timeout=15, max=100 Location: http://domain.com/ [following] And a working one: user@localhost dir $ wget -S domain.com --2010-03-20 21:51:33-- http://domain.com/ Resolving domain.com... x.x.x.x Connecting to domain.com|x.x.x.x|:80... connected. HTTP request sent, awaiting response... HTTP/1.1 301 Moved Permanently Via: 1.1 ISA Connection: Keep-Alive Proxy-Connection: Keep-Alive Content-Length: 0 Date: Sat, 20 Mar 2010 20:51:33 GMT Location: http://www.domain.com/ Content-Type: text/html; charset=UTF-8 Server: Apache X-Powered-By: PHP/5.2.10-pl0-gentoo X-Pingback: http://www.domain.com/xmlrpc.php Keep-Alive: timeout=15, max=100 Location: http://www.domain.com/ [following] I'm stumped. I've had this problem for a long time, and I feel like I've tried everything. I can't see why the different domains would act differently under the same installation with the same settings. Help :(

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  • Exchange Server 2007 Setup

    - by AlamedaDad
    Hi, I'm working on a upgrade to Exchange 2007 and I wanted to get some advise on hardware choices. We currently have an Exchange 2003 STD server with 400 users split between 6 AD Sites, that is housed on a single server. We need to move to a redundant, fault tolerant system to support our users. I'm planning on installing 2 Dell 1950 servers with W2k8-std to act as CAS and Hub servers, with NLB to allow abstraction of the actual server name to the users. There won't be an edge system since we have a Barracuda box already that will handle in/out spam/virus filtering. Backend I'm planning on 2 mailbox servers which will be Dell 2950s with 16GB RAM, 2 either dual-core or quad-core CPUs and 6 300GB SAS drives in some RAID config. These systems will be clustered using W2k8 Ent clustering and running CCR in Exchange. My questions are as follows: Is 16GB enough RAM for serving that many mailboxes along with the windows clustering and ccr? I'm trying to figure out disk layouts and I'm unsure of whether to use all local disk or some local and some SAN, via an OpenFiler iSCSI server. The SAN would be a Dell 2850 with 6 - 300GB SCSI drives and a PERC controller to slice as I want, with 8GB RAM. Option 1: 2 drives, RAID 1 - OS 2 drives, RAID 1 - Logs 2 drives, RAID 1 - Mail stores Option 2: 2 drives, RAID 1 - OS and logs 4 drives, RAID 5 - Mail Stores and scratch space for eseutil. Option 3: 2 drives, RAID 1 - OS 2 drives, RAID 1 - Logs 2 drives, RAID 0 - scratch space ~300GB iSCSI volume for mail stores Option 4: 2 drives, RAID 1 - OS 4 drives, RAID 5 - scratch space ~300GB iSCSI volume for mail stores ~300GB iSCSI volume for logs I have 2 sockets for CPUs and need to chose between dual and quad cores. The dual core have faster clocks but less cache and I'm thinking older architecture. Am I better off with more cores and cache while sacraficing clock speed? I am planning on adding the new E2K7 cluster to the E2K3 server and then move each mailbox over, all at once, then remove the old server. This seems more complicated than simply getting rid of the 2003 server and then adding the 2007 cluster and restoring the mailboxes using PowerControls or exmerge. The migration option lets me do this on my time, where a cutover means it all needs to work at once. If I go with the cutover method, how can I prebuild the servers and add them to the domain right after removing the 2003 server, or can't I? I think the answer is no and the migration is my only real option if I want to prebuild. I need to also migrate about 30GB of Public Folders. Is there anything special about this, other than specifying in the E2K7 install that I want older Outlook clients and PF's setup? I guess I could even keep the E2K3 server to host just the PFs? Lastly, if I have a mix of Outlook 200, 2003 and 2007 what do I need to do to make sure they all have access to the GAL and OAB? At time of cutover, we'll be at like 90% 2007, but we will have some older stuff around. My plan is to use Outlook Anywhere on laptops that are used outside the physical network. Are there any gotchas involved in that? I'm even thinking about using is for all Outlook clients, does anyone do that? The reason I'm considering it is that our WAN is really VPN tunnels over internet connections, so not a fully messhed, stable WAN. Thank you all very much for the assistance in advance and I look forward to discussion of these points! Regards...Michael

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  • Windows 7 BSOD - ntoskrnl?

    - by Ken Mason
    2 new HP Pavilion notebooks with 7 Home Premium pre-loaded with Norton. My first act was to use the Norton Removal Tool and load ZoneAlarm free and AVG Free. Frequent random BSOD's ever since...I found my way into Debug and have had various reports regarding ntoskrnl, depending on the status of symbols. It's been many years since I played with (DOS 3.x) debug, so this has been a considerable fumble. Excerpts follow and any insights would be greatly appreciated, as I am not a developer: ADDITIONAL_DEBUG_TEXT: Use '!findthebuild' command to search for the target build information. If the build information is available, run '!findthebuild -s ; .reload' to set symbol path and load symbols. MODULE_NAME: nt FAULTING_MODULE: fffff8000305d000 nt DEBUG_FLR_IMAGE_TIMESTAMP: 4b88cfeb BUGCHECK_STR: 0x7f_8 CUSTOMER_CRASH_COUNT: 1 DEFAULT_BUCKET_ID: VISTA_DRIVER_FAULT CURRENT_IRQL: 0 LAST_CONTROL_TRANSFER: from fffff800030ccb69 to fffff800030cd600 STACK_TEXT: fffff80004d6fd28 fffff800030ccb69 : 000000000000007f 0000000000000008 0000000080050033 00000000000006f8 : nt+0x70600 fffff80004d6fd30 000000000000007f : 0000000000000008 0000000080050033 00000000000006f8 fffff80003095e58 : nt+0x6fb69 fffff80004d6fd38 0000000000000008 : 0000000080050033 00000000000006f8 fffff80003095e58 0000000000000000 : 0x7f fffff80004d6fd40 0000000080050033 : 00000000000006f8 fffff80003095e58 0000000000000000 0000000000000000 : 0x8 fffff80004d6fd48 00000000000006f8 : fffff80003095e58 0000000000000000 0000000000000000 0000000000000000 : 0x80050033 fffff80004d6fd50 fffff80003095e58 : 0000000000000000 0000000000000000 0000000000000000 0000000000000000 : 0x6f8 fffff80004d6fd58 0000000000000000 : 0000000000000000 0000000000000000 0000000000000000 0000000000000000 : nt+0x38e58 STACK_COMMAND: kb FOLLOWUP_IP: nt+70600 fffff800`030cd600 48894c2408 mov qword ptr [rsp+8],rcx SYMBOL_STACK_INDEX: 0 SYMBOL_NAME: nt+70600 FOLLOWUP_NAME: MachineOwner IMAGE_NAME: ntoskrnl.exe BUCKET_ID: WRONG_SYMBOLS Followup: MachineOwner ...................................................................... 0: kd !lmi nt Loaded Module Info: [nt] Module: ntkrnlmp Base Address: fffff8000305d000 Image Name: ntkrnlmp.exe Machine Type: 34404 (X64) Time Stamp: 4b88cfeb Sat Feb 27 00:55:23 2010 Size: 5dc000 CheckSum: 545094 Characteristics: 22 perf Debug Data Dirs: Type Size VA Pointer CODEVIEW 25, 19c65c, 19bc5c RSDS - GUID: {7E9A3CAB-6268-45DE-8E10-816E3080A3B7} Age: 2, Pdb: ntkrnlmp.pdb CLSID 4, 19c658, 19bc58 [Data not mapped] Image Type: FILE - Image read successfully from debugger. ntkrnlmp.exe Symbol Type: PDB - Symbols loaded successfully from symbol server. d:\debugsymbols\ntkrnlmp.pdb\7E9A3CAB626845DE8E10816E3080A3B72\ntkrnlmp.pdb Load Report: public symbols , not source indexed d:\debugsymbols\ntkrnlmp.pdb\7E9A3CAB626845DE8E10816E3080A3B72\ntkrnlmp.pdb 0: kd !analyze -v * Bugcheck Analysis * * UNEXPECTED_KERNEL_MODE_TRAP (7f) This means a trap occurred in kernel mode, and it's a trap of a kind that the kernel isn't allowed to have/catch (bound trap) or that is always instant death (double fault). The first number in the bugcheck params is the number of the trap (8 = double fault, etc) Consult an Intel x86 family manual to learn more about what these traps are. Here is a portion of those codes: If kv shows a taskGate use .tss on the part before the colon, then kv. Else if kv shows a trapframe use .trap on that value Else .trap on the appropriate frame will show where the trap was taken (on x86, this will be the ebp that goes with the procedure KiTrap) Endif kb will then show the corrected stack. Arguments: Arg1: 0000000000000008, EXCEPTION_DOUBLE_FAULT Arg2: 0000000080050033 Arg3: 00000000000006f8 Arg4: fffff80003095e58 Debugging Details: BUGCHECK_STR: 0x7f_8 CUSTOMER_CRASH_COUNT: 1 DEFAULT_BUCKET_ID: VISTA_DRIVER_FAULT PROCESS_NAME: System CURRENT_IRQL: 2 LAST_CONTROL_TRANSFER: from fffff800030ccb69 to fffff800030cd600 STACK_TEXT: fffff80004d6fd28 fffff800030ccb69 : 000000000000007f 0000000000000008 0000000080050033 00000000000006f8 : nt!KeBugCheckEx fffff80004d6fd30 fffff800030cb032 : 0000000000000000 0000000000000000 0000000000000000 0000000000000000 : nt!KiBugCheckDispatch+0x69 fffff80004d6fe70 fffff80003095e58 : 0000000000000000 0000000000000000 0000000000000000 0000000000000000 : nt!KiDoubleFaultAbort+0xb2 fffff880089efc60 0000000000000000 : 0000000000000000 0000000000000000 0000000000000000 0000000000000000 : nt!SeAccessCheckFromState+0x58 STACK_COMMAND: kb FOLLOWUP_IP: nt!KiDoubleFaultAbort+b2 fffff800`030cb032 90 nop SYMBOL_STACK_INDEX: 2 SYMBOL_NAME: nt!KiDoubleFaultAbort+b2 FOLLOWUP_NAME: MachineOwner MODULE_NAME: nt IMAGE_NAME: ntkrnlmp.exe DEBUG_FLR_IMAGE_TIMESTAMP: 4b88cfeb FAILURE_BUCKET_ID: X64_0x7f_8_nt!KiDoubleFaultAbort+b2 BUCKET_ID: X64_0x7f_8_nt!KiDoubleFaultAbort+b2 Followup: MachineOwner I tried running Rootkit Revealer but I don't think it works on x64 systems. Similarly Blacklight seems to have aged off. I'm running Sophos Anti-Rootkit now. So far so good...

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  • Blocking 'good' bots in nginx with multiple conditions for certain off-limits URL's where humans can go

    - by Glenn Plas
    After 2 days of searching/trying/failing I decided to post this here, I haven't found any example of someone doing the same nor what I tried seems to be working OK. I'm trying to send a 403 to bots not respecting the robots.txt file (even after downloading it several times). Specifically Googlebot. It will support the following robots.txt definition. User-agent: * Disallow: /*/*/page/ The intent is to allow Google to browse whatever they can find on the site but return a 403 for the following type of request. Googlebot seems to keep on nesting these links eternally adding paging block after block: my_domain.com:80 - 66.x.67.x - - [25/Apr/2012:11:13:54 +0200] "GET /2011/06/ page/3/?/page/2//page/3//page/2//page/3//page/2//page/2//page/4//page/4//pag e/1/&wpmp_switcher=desktop HTTP/1.1" 403 135 "-" "Mozilla/5.0 (compatible; G ooglebot/2.1; +http://www.google.com/bot.html)" It's a wordpress site btw. I don't want those pages to show up, even though after the robots.txt info got through, they stopped for a while only to begin crawling again later. It just never stops .... I do want real people to see this. As you can see, google get a 403 but when I try this myself in a browser I get a 404 back. I want browsers to pass. root@my_domain:# nginx -V nginx version: nginx/1.2.0 I tried different approaches, using a map and plain old nono if's and they both act the same: (under http section) map $http_user_agent $is_bot { default 0; ~crawl|Googlebot|Slurp|spider|bingbot|tracker|click|parser|spider 1; } (under the server section) location ~ /(\d+)/(\d+)/page/ { if ($is_bot) { return 403; # Please respect the robots.txt file ! } } I recently had to polish up my Apache skills for a client where I did about the same thing like this : # Block real Engines , not respecting robots.txt but allowing correct calls to pass # Google RewriteCond %{HTTP_USER_AGENT} ^Mozilla/5\.0\ \(compatible;\ Googlebot/2\.[01];\ \+http://www\.google\.com/bot\.html\)$ [NC,OR] # Bing RewriteCond %{HTTP_USER_AGENT} ^Mozilla/5\.0\ \(compatible;\ bingbot/2\.[01];\ \+http://www\.bing\.com/bingbot\.htm\)$ [NC,OR] # msnbot RewriteCond %{HTTP_USER_AGENT} ^msnbot-media/1\.[01]\ \(\+http://search\.msn\.com/msnbot\.htm\)$ [NC,OR] # Slurp RewriteCond %{HTTP_USER_AGENT} ^Mozilla/5\.0\ \(compatible;\ Yahoo!\ Slurp;\ http://help\.yahoo\.com/help/us/ysearch/slurp\)$ [NC] # block all page searches, the rest may pass RewriteCond %{REQUEST_URI} ^(/[0-9]{4}/[0-9]{2}/page/) [OR] # or with the wpmp_switcher=mobile parameter set RewriteCond %{QUERY_STRING} wpmp_switcher=mobile # ISSUE 403 / SERVE ERRORDOCUMENT RewriteRule .* - [F,L] # End if match This does a bit more than I asked nginx to do but it's about the same principle, I'm having a hard time figuring this out for nginx. So my question would be, why would nginx serve my browser a 404 ? Why isn't it passing, The regex isn't matching for my UA: "Mozilla/5.0 (X11; Linux x86_64) AppleWebKit/536.5 (KHTML, like Gecko) Chrome/19.0.1084.30 Safari/536.5" There are tons of example to block based on UA alone, and that's easy. It also looks like the matchin location is final, e.g. it's not 'falling' through for regular user, I'm pretty certain that this has some correlation with the 404 I get in the browser. As a cherry on top of things, I also want google to disregard the parameter wpmp_switcher=mobile , wpmp_switcher=desktop is fine but I just don't want the same content being crawled multiple times. Even though I ended up adding wpmp_switcher=mobile via the google webmaster tools pages (requiring me to sign up ....). that also stopped for a while but today they are back spidering the mobile sections. So in short, I need to find a way for nginx to enforce the robots.txt definitions. Can someone shell out a few minutes of their lives and push me in the right direction please ? I really appreciate ANY response that makes me think harder ;-)

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  • Just a few questions about Hyper-V virtual machines and clustering

    - by René Kåbis
    I have been using Microsoft’s Hyper-V technology for a little while now, but I am just now dipping my toe into clustering. In particular, I am trying to implement a fault-tolerant SQL DB. This involves setting up two VMs, clustering them via Failover Cluster, and then installing SQL Server in some fashion. I have two physical machines - one high-end and rather beefy “heavy lifter” to contain the majority of the VMs, and another “backup” (a repurposed desktop) to hold the essential “secondary” (or failover) AD-DC, SQL and FS VMs. The main reason why I find the failover cluster at the VM level so attractive is that it presents a single IP and DNS entry to the network as a whole - if one machine (physical or virtual) goes down, you might loose some ping and the connections get reset, but the network applications (Microsoft RMS connection to backend SQL) can still connect to a viable DB without having to mess around with the settings at all. My first question is in terms of SQL Server itself. If I have a cluster between two VMs, does it make more sense to install the SQL Server in Failover Cluster configuration or should I simply install it in a stand-alone config and mirror the DBs? For example, this post suggests just mirroring the DBs, but do I just mirror standalone DBs on standalone VMs, or can I get the network and failover benefits of clustered VMs while still utilizing (on each clustered VM) standalone DBs that have been mirrored between each other? As well, I have come across a lot of documentation about SQL clustering, but most assume a number (#2) of physical machines to hold not only the actual SQL VMs but also the Quorum and Witness stores. I will not be able to muster more than two physical machines. As such, I will have to be satisfied with a VM cluster that does not exceed two VMs (one for each physical machine). Another issue involves MSDTC - the Distributed Transaction Coordinator. When attempting to install the SQL Failover Cluster (I never completed it for this reason) it threw a hissy fit because MSDTC had not been clustered. Search as I might, I have not yet found a way to do so under Windows Server 2012 R2. I have found plenty of docs for Windows 2008 and 2008 R2, but these instructions don’t align with 2012 R2 (at least, not in a way that allows me to successfully cluster MSDTC). Plus, some of the instructions that I have found for SQL Server Failover Cluster installation suggest that a third “network device” - shared network storage (a SAN) - is required for the DB itself (and other functionality). I do not have this, and won’t be getting this. Most of my storage exists on the “heavy lifter” that was designed for all of the “primary” VMs. If that physical machine goes down, so does the storage. The secondary server does have enough resources for an AD-DC Server, an SQL server and a File Server, so it will handle the “secondary” failover versions of those VMs (clustered or not). My final question involves file servers. If I cluster file servers between two VMs (one on my “heavy lifter” and another on my “backup”, how do I mirror the data between them? Clustering VMs only provides a single point of access on the network for a resource, it doesn’t exactly replicate data between the two - that is left to the services that serve up that data. I am unsure how I can ensure that file server data between two clustered file server VMs can be properly mirrored. Remember, I only have two devices to be used here - my primary machine and a backup secondary. There is no chance of me obtaining a SAN or any other type of network attached storage. What exists on the machines must act as the storage. Thanks in advance for any suggestions.

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  • Parallel processing slower than sequential?

    - by zebediah49
    EDIT: For anyone who stumbles upon this in the future: Imagemagick uses a MP library. It's faster to use available cores if they're around, but if you have parallel jobs, it's unhelpful. Do one of the following: do your jobs serially (with Imagemagick in parallel mode) set MAGICK_THREAD_LIMIT=1 for your invocation of the imagemagick binary in question. By making Imagemagick use only one thread, it slows down by 20-30% in my test cases, but meant I could run one job per core without issues, for a significant net increase in performance. Original question: While converting some images using ImageMagick, I noticed a somewhat strange effect. Using xargs was significantly slower than a standard for loop. Since xargs limited to a single process should act like a for loop, I tested that, and found it to be about the same. Thus, we have this demonstration. Quad core (AMD Athalon X4, 2.6GHz) Working entirely on a tempfs (16g ram total; no swap) No other major loads Results: /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 1 convert -auto-level real 0m3.784s user 0m2.240s sys 0m0.230s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 2 convert -auto-level real 0m9.097s user 0m28.020s sys 0m0.910s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 10 convert -auto-level real 0m9.844s user 0m33.200s sys 0m1.270s Can anyone think of a reason why running two instances of this program takes more than twice as long in real time, and more than ten times as long in processor time to complete the same task? After that initial hit, more processes do not seem to have as significant of an effect. I thought it might have to do with disk seeking, so I did that test entirely in ram. Could it have something to do with how Convert works, and having more than one copy at once means it cannot use processor cache as efficiently or something? EDIT: When done with 1000x 769KB files, performance is as expected. Interesting. /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 1 convert -auto-level real 3m37.679s user 5m6.980s sys 0m6.340s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 1 convert -auto-level real 3m37.152s user 5m6.140s sys 0m6.530s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 2 convert -auto-level real 2m7.578s user 5m35.410s sys 0m6.050s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 4 convert -auto-level real 1m36.959s user 5m48.900s sys 0m6.350s /media/ramdisk/img$ time for f in *.bmp; do echo $f ${f%bmp}png; done | xargs -n 2 -P 10 convert -auto-level real 1m36.392s user 5m54.840s sys 0m5.650s

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  • volume group disappeared after xfs_check run

    - by John P
    EDIT** I have a volume group consisting of 5 RAID1 devices grouped together into a lvm and formatted with xfs. The 5th RAID device lost its RAID config (cat /proc/mdstat does not show anything). The two drives are still present (sdj and sdk), but they have no partitions. The LVM appeared to be happily using sdj up until recently. (doing a pvscan showed the first 4 RAID1 devices + /dev/sdj) I removed the LVM from the fstab, rebooted, then ran xfs_check on the LV. It ran for about half an hour, then stopped with an error. I tried rebooting again, and this time when it came up, the logical volume was no longer there. It is now looking for /dev/md5, which is gone (though it had been using /dev/sdj earlier). /dev/sdj was having read errors, but after replacing the SATA cable, those went away, so the drive appears to be fine for now. Can I modify the /etc/lvm/backup/dedvol, change the device to /dev/sdj and do a vgcfgrestore? I could try doing a pvcreate --uuid KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ /dev/sdj to make it recognize it, but I'm afraid that would erase the data on the drive UPDATE: just changing the pv to point to /dev/sdj did not work vgcfgrestore --file /etc/lvm/backup/dedvol dedvol Couldn't find device with uuid 'KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ'. Cannot restore Volume Group dedvol with 1 PVs marked as missing. Restore failed. pvscan /dev/sdj: read failed after 0 of 4096 at 0: Input/output error Couldn't find device with uuid 'KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ'. Couldn't find device with uuid 'KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ'. Couldn't find device with uuid 'KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ'. Couldn't find device with uuid 'KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ'. PV /dev/sdd2 VG VolGroup00 lvm2 [74.41 GB / 0 free] PV /dev/md2 VG dedvol lvm2 [931.51 GB / 0 free] PV /dev/md3 VG dedvol lvm2 [931.51 GB / 0 free] PV /dev/md0 VG dedvol lvm2 [931.51 GB / 0 free] PV /dev/md4 VG dedvol lvm2 [931.51 GB / 0 free] PV unknown device VG dedvol lvm2 [1.82 TB / 63.05 GB free] Total: 6 [5.53 TB] / in use: 6 [5.53 TB] / in no VG: 0 [0 ] vgscan Reading all physical volumes. This may take a while... /dev/sdj: read failed after 0 of 4096 at 0: Input/output error /dev/sdj: read failed after 0 of 4096 at 2000398843904: Input/output error Found volume group "VolGroup00" using metadata type lvm2 Found volume group "dedvol" using metadata type lvm2 vgdisplay dedvol --- Volume group --- VG Name dedvol System ID Format lvm2 Metadata Areas 5 Metadata Sequence No 10 VG Access read/write VG Status resizable MAX LV 0 Cur LV 1 Open LV 0 Max PV 0 Cur PV 5 Act PV 5 VG Size 5.46 TB PE Size 4.00 MB Total PE 1430796 Alloc PE / Size 1414656 / 5.40 TB Free PE / Size 16140 / 63.05 GB VG UUID o1U6Ll-5WH8-Pv7Z-Rtc4-1qYp-oiWA-cPD246 dedvol { id = "o1U6Ll-5WH8-Pv7Z-Rtc4-1qYp-oiWA-cPD246" seqno = 10 status = ["RESIZEABLE", "READ", "WRITE"] flags = [] extent_size = 8192 # 4 Megabytes max_lv = 0 max_pv = 0 physical_volumes { pv0 { id = "Msiee7-Zovu-VSJ3-Y2hR-uBVd-6PaT-Ho9v95" device = "/dev/md2" # Hint only status = ["ALLOCATABLE"] flags = [] dev_size = 1953519872 # 931.511 Gigabytes pe_start = 384 pe_count = 238466 # 931.508 Gigabytes } pv1 { id = "ZittCN-0x6L-cOsW-v1v4-atVN-fEWF-e3lqUe" device = "/dev/md3" # Hint only status = ["ALLOCATABLE"] flags = [] dev_size = 1953519872 # 931.511 Gigabytes pe_start = 384 pe_count = 238466 # 931.508 Gigabytes } pv2 { id = "NRNo0w-kgGr-dUxA-mWnl-bU5v-Wld0-XeKVLD" device = "/dev/md0" # Hint only status = ["ALLOCATABLE"] flags = [] dev_size = 1953519872 # 931.511 Gigabytes pe_start = 384 pe_count = 238466 # 931.508 Gigabytes } pv3 { id = "2EfLFr-JcRe-MusW-mfAs-WCct-u4iV-W0pmG3" device = "/dev/md4" # Hint only status = ["ALLOCATABLE"] flags = [] dev_size = 1953519872 # 931.511 Gigabytes pe_start = 384 pe_count = 238466 # 931.508 Gigabytes } pv4 { id = "KZron2-pPTr-ZYeQ-PKXX-4Woq-6aNc-AG4rRJ" device = "/dev/md5" # Hint only status = ["ALLOCATABLE"] flags = [] dev_size = 3907028992 # 1.81935 Terabytes pe_start = 384 pe_count = 476932 # 1.81935 Terabytes } }

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  • How do I make an external hard drive keep the same drive letter permanently?

    - by andygrunt
    I have a desktop PC (2002 vintage) running Windows XP that I turn on about 2 or 3 times per week. I have a mains powered 250Gb Western Digital hard disk connected to it via USB. I always turn the hard disk on before the PC so it's up and running as the PC boots. When I first connected the external hard disk, the PC assigned it a letter ('i' if it matters) and I've installed software to it, created shortcuts to various files and folders on the disk using that letter. For years everything was fine then I would boot the PC and the hard disk was assigned a different letter. I'd then have to go into 'my computer/manage/disk management' and manually change the letter back to 'i'. If I then rebooted the PC, the hard disk would usually still be 'i' but after the next reboot would be some other random letter and I have to manually change it back to 'i'. This would go on for some time then there'd be periods when the it would always be 'i' then, for no apparent reason (no new devices added, for example), the drive letter would start changing again. At the moment it's in random drive letter mood so I thought I'd ask the following question... How do I assign the external hard disk to be 'i' permanently? Answer: Thanks Molly that seems to have done the trick (after a little fiddling) - slightly disappointed there wasn't a way to do it within Windows without installing something else though. For anyone else trying this, it wasn't completely straightforward so here's what happened with me. I installed USBDLM as per the instructions on its website. I guessed that I had to assign the first USB letter to i so replaced the 'Letter1=' lines to 'Letter=I' in the ini file. To test it, I rebooted the PC only to find it came back up with the display set to 640x480 in 16 colours. After some investigation, I re-installed the display drivers and rebooted and set the display back to its usual setting. The external hard disk now gets set to 'i' but I found I had to re-apply sharing status to it so it was seen from my laptop which is on the same network. The end result of all this is that it now does what I wanted although it does act as though the hard drive has just been plugged in a few seconds after the Windows desktop appears i.e. the little box appears with a progress bar as it searches through the contents of the 'new' hard drive and I eventually get a dialogue box saying 'This disk or device contains more than one type of content. What do you want Windows to do?' and lists options such as play media files, print the pictures or open folder to view the files. This is a tiny pain I wish didn't happen but not exactly a huge price to pay. Other than that - it seems to work fine :) Looks like a spoke too soon... Every time I reboot, I have to re-share the 'i' drive (which I didn't have to do before) so it can be seen by my laptop on the same network. Any ideas how to make that permanent?

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  • How does the Cloud compare to Colocation? And development too

    - by David
    Currently I/we run a SaaS web application where each subscriber has their own physical instance of the application in addition to their own database. The setup has each web application instance deployed on two different IIS boxes both for load-balancing and redundancy (the machines have their Windows Update install times 12 hours apart, for example). Databases are mirrored on two different SQL Server 2012 machines with AlwaysOn for uptime. I don't make use of SQL Server clustering (as it doesn't provide storage-level failover: we don't have a shared storage box). Because it's a Windows setup it means there are two Domain Controllers (we cheat: they're both Mac Minis, 17W each, which keeps our colo power costs low). Finally there's also an Exchange server (Mailbox, Hub Transport and Client Access). One of the SQL Servers also doubles-up as an Exchange Hub Transport. Running costs are about $700 a month for our quarter-rack colocation (which includes power and peering/transfer), then there's about $150 a month for SPLA licensing, so $850 a month in total. Then there's the hard-to-quantify cost of administration, but I reckon I spend a couple of hours a week checking-in on the servers: reviewing event logs, etc. I keep getting bombarded by ads and manufactured news stories about how great "the cloud" is. Back in 2008 when the cloud was taking off I was reading up about the proper "cloud" services like Google AppEngine, where you write in Python against Google's API and that's how they scale your application across servers and also use their database provider for scaling storage. Simple enough to understand. Then came along Amazon, and I understand how Amazon Storage works, but I'm not sure how Amazon Compute works: web application pages don't take much CPU time to compute, how do you even quantify usage anyway? Finally, RackSpace gets in the act and now I'm really confused. RackSpace advertise "Cloud" SQL Server 2012 available for about "$0.70 per hour", going by how they advertise it I thought the "hour" meant the sum of CPU time, IO blocking time, maybe time spent transferring data, so for a low-intensity application that works out pretty cheap then? Nope. I went on to a Sales Chat window and spoke to one of their advisors. They told me the $0.70/hour was actually for every hour the SQL Server is running... but who wants a SQL Server for only a few hours? You're going to need it available 24 hours a day for months on end. $0.70 * 24 * 31 works out at $520 a month, which is rediculously expensive for SQL Server. An SPLA license for SQL Server is only $50 a month or so. That $520 a month does not include "fanatical support", and you also need to stack on top the costs of the host Windows server instance too. From what I can tell, Rackspace's "Cloud" products seem like like an cynical rebranding of an overpriced VPS service, but priced by the hour. I have the same confusion about Windows Azure which uses similar terms to describe the products available, but I think that's because Azure offers both traditional shared webhosting in addition to their own APIs you can target for scalable applications.

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  • Behind ASP.NET MVC Mock Objects

    - by imran_ku07
       Introduction:           I think this sentence now become very familiar to ASP.NET MVC developers that "ASP.NET MVC is designed with testability in mind". But what ASP.NET MVC team did for making applications build with ASP.NET MVC become easily testable? Understanding this is also very important because it gives you some help when designing custom classes. So in this article i will discuss some abstract classes provided by ASP.NET MVC team for the various ASP.NET intrinsic objects, including HttpContext, HttpRequest, and HttpResponse for making these objects as testable. I will also discuss that why it is hard and difficult to test ASP.NET Web Forms.      Description:           Starting from Classic ASP to ASP.NET MVC, ASP.NET Intrinsic objects is extensively used in all form of web application. They provide information about Request, Response, Server, Application and so on. But ASP.NET MVC uses these intrinsic objects in some abstract manner. The reason for this abstraction is to make your application testable. So let see the abstraction.           As we know that ASP.NET MVC uses the same runtime engine as ASP.NET Web Form uses, therefore the first receiver of the request after IIS and aspnet_filter.dll is aspnet_isapi.dll. This will start the application domain. With the application domain up and running, ASP.NET does some initialization and after some initialization it will call Application_Start if it is defined. Then the normal HTTP pipeline event handlers will be executed including both HTTP Modules and global.asax event handlers. One of the HTTP Module is registered by ASP.NET MVC is UrlRoutingModule. The purpose of this module is to match a route defined in global.asax. Every matched route must have IRouteHandler. In default case this is MvcRouteHandler which is responsible for determining the HTTP Handler which returns MvcHandler (which is derived from IHttpHandler). In simple words, Route has MvcRouteHandler which returns MvcHandler which is the IHttpHandler of current request. In between HTTP pipeline events the handler of ASP.NET MVC, MvcHandler.ProcessRequest will be executed and shown as given below,          void IHttpHandler.ProcessRequest(HttpContext context)          {                    this.ProcessRequest(context);          }          protected virtual void ProcessRequest(HttpContext context)          {                    // HttpContextWrapper inherits from HttpContextBase                    HttpContextBase ctxBase = new HttpContextWrapper(context);                    this.ProcessRequest(ctxBase);          }          protected internal virtual void ProcessRequest(HttpContextBase ctxBase)          {                    . . .          }             HttpContextBase is the base class. HttpContextWrapper inherits from HttpContextBase, which is the parent class that include information about a single HTTP request. This is what ASP.NET MVC team did, just wrap old instrinsic HttpContext into HttpContextWrapper object and provide opportunity for other framework to provide their own implementation of HttpContextBase. For example           public class MockHttpContext : HttpContextBase          {                    . . .          }                     As you can see, it is very easy to create your own HttpContext. That's what did the third party mock frameworks like TypeMock, Moq, RhinoMocks, or NMock2 to provide their own implementation of ASP.NET instrinsic objects classes.           The key point to note here is the types of ASP.NET instrinsic objects. In ASP.NET Web Form and ASP.NET MVC. For example in ASP.NET Web Form the type of Request object is HttpRequest (which is sealed) and in ASP.NET MVC the type of Request object is HttpRequestBase. This is one of the reason that makes test in ASP.NET WebForm is difficult. because their is no base class and the HttpRequest class is sealed, therefore it cannot act as a base class to others. On the other side ASP.NET MVC always uses a base class to give a chance to third parties and unit test frameworks to create thier own implementation ASP.NET instrinsic object.           Therefore we can say that in ASP.NET MVC, instrinsic objects are of type base classes (for example HttpContextBase) .Actually these base classes had it's own implementation of same interface as the intrinsic objects it abstracts. It includes only virtual members which simply throws an exception. ASP.NET MVC also provides the corresponding wrapper classes (for example, HttpRequestWrapper) which provides a concrete implementation of the base classes in the form of ASP.NET intrinsic object. Other wrapper classes may be defined by third parties in the form of a mock object for testing purpose.           So we can say that a Request object in ASP.NET MVC may be HttpRequestWrapper or may be MockRequestWrapper(assuming that MockRequestWrapper class is used for testing purpose). Here is list of ASP.NET instrinsic and their implementation in ASP.NET MVC in the form of base and wrapper classes. Base Class Wrapper Class ASP.NET Intrinsic Object Description HttpApplicationStateBase HttpApplicationStateWrapper Application HttpApplicationStateBase abstracts the intrinsic Application object HttpBrowserCapabilitiesBase HttpBrowserCapabilitiesWrapper HttpBrowserCapabilities HttpBrowserCapabilitiesBase abstracts the HttpBrowserCapabilities class HttpCachePolicyBase HttpCachePolicyWrapper HttpCachePolicy HttpCachePolicyBase abstracts the HttpCachePolicy class HttpContextBase HttpContextWrapper HttpContext HttpContextBase abstracts the intrinsic HttpContext object HttpFileCollectionBase HttpFileCollectionWrapper HttpFileCollection HttpFileCollectionBase abstracts the HttpFileCollection class HttpPostedFileBase HttpPostedFileWrapper HttpPostedFile HttpPostedFileBase abstracts the HttpPostedFile class HttpRequestBase HttpRequestWrapper Request HttpRequestBase abstracts the intrinsic Request object HttpResponseBase HttpResponseWrapper Response HttpResponseBase abstracts the intrinsic Response object HttpServerUtilityBase HttpServerUtilityWrapper Server HttpServerUtilityBase abstracts the intrinsic Server object HttpSessionStateBase HttpSessionStateWrapper Session HttpSessionStateBase abstracts the intrinsic Session object HttpStaticObjectsCollectionBase HttpStaticObjectsCollectionWrapper HttpStaticObjectsCollection HttpStaticObjectsCollectionBase abstracts the HttpStaticObjectsCollection class      Summary:           ASP.NET MVC provides a set of abstract classes for ASP.NET instrinsic objects in the form of base classes, allowing someone to create their own implementation. In addition, ASP.NET MVC also provide set of concrete classes in the form of wrapper classes. This design really makes application easier to test and even application may replace concrete implementation with thier own implementation, which makes ASP.NET MVC very flexable.

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  • TemplateBinding with Converter - what is wrong?

    - by MartyIX
    I'm creating a game desk. I wanted to specify field size (one field is a square) as a attached property and with this data set value of ViewPort which would draw 2x2 matrix (and tile mode would do the rest of game desk). I'm quite at loss what is wrong because the binding doesn't work. Testing line in XAML for the behaviour I would like to have: <DrawingBrush Viewport="0,0,100,100" ViewportUnits="Absolute" TileMode="None"> The game desk is based on this sample of DrawingPaint: http://msdn.microsoft.com/en-us/library/aa970904.aspx (an image is here) XAML: <Window x:Class="Sokoban.Window1" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:local="clr-namespace:Sokoban" Title="Window1" Height="559" Width="419"> <Window.Resources> <local:FieldSizeToRectConverter x:Key="fieldSizeConverter" /> <Style x:Key="GameDesk" TargetType="{x:Type Rectangle}"> <Setter Property="local:GameDeskProperties.FieldSize" Value="50" /> <Setter Property="Fill"> <Setter.Value> <!--<DrawingBrush Viewport="0,0,100,100" ViewportUnits="Absolute" TileMode="None">--> <DrawingBrush Viewport="{TemplateBinding local:GameDeskProperties.FieldSize, Converter={StaticResource fieldSizeConverter}}" ViewportUnits="Absolute" TileMode="None"> <DrawingBrush.Drawing> <DrawingGroup> <GeometryDrawing Brush="CornflowerBlue"> <GeometryDrawing.Geometry> <RectangleGeometry Rect="0,0,100,100" /> </GeometryDrawing.Geometry> </GeometryDrawing> <GeometryDrawing Brush="Azure"> <GeometryDrawing.Geometry> <GeometryGroup> <RectangleGeometry Rect="0,0,50,50" /> <RectangleGeometry Rect="50,50,50,50" /> </GeometryGroup> </GeometryDrawing.Geometry> </GeometryDrawing> </DrawingGroup> </DrawingBrush.Drawing> </DrawingBrush> </Setter.Value> </Setter> </Style> </Window.Resources> <StackPanel> <Rectangle Style="{StaticResource GameDesk}" Width="300" Height="150" /> </StackPanel> </Window> Converter and property definition: using System; using System.Collections.Generic; using System.Text; using System.Windows.Controls; using System.Windows; using System.Diagnostics; using System.Windows.Data; namespace Sokoban { public class GameDeskProperties : Panel { public static readonly DependencyProperty FieldSizeProperty; static GameDeskProperties() { PropertyChangedCallback fieldSizeChanged = new PropertyChangedCallback(OnFieldSizeChanged); PropertyMetadata fieldSizeMetadata = new PropertyMetadata(50, fieldSizeChanged); FieldSizeProperty = DependencyProperty.RegisterAttached("FieldSize", typeof(int), typeof(GameDeskProperties), fieldSizeMetadata); } public static int GetFieldSize(DependencyObject target) { return (int)target.GetValue(FieldSizeProperty); } public static void SetFieldSize(DependencyObject target, int value) { target.SetValue(FieldSizeProperty, value); } static void OnFieldSizeChanged(DependencyObject target, DependencyPropertyChangedEventArgs e) { Debug.WriteLine("FieldSize just changed: " + e.NewValue); } } [ValueConversion(/* sourceType */ typeof(int), /* targetType */ typeof(Rect))] public class FieldSizeToRectConverter : IValueConverter { public object Convert(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture) { Debug.Assert(targetType == typeof(int)); int fieldSize = int.Parse(value.ToString()); return new Rect(0, 0, 2 * fieldSize, 2 * fieldSize); } public object ConvertBack(object value, Type targetType, object parameter, System.Globalization.CultureInfo culture) { // should not be called in our example throw new NotImplementedException(); } } }

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  • jQuery Time Entry with Time Navigation Keys

    - by Rick Strahl
    So, how do you display time values in your Web applications? Displaying date AND time values in applications is lot less standardized than date display only. While date input has become fairly universal with various date picker controls available, time entry continues to be a bit of a non-standardized. In my own applications I tend to use the jQuery UI DatePicker control for date entries and it works well for that. Here's an example: The date entry portion is well defined and it makes perfect sense to have a calendar pop up so you can pick a date from a rich UI when necessary. However, time values are much less obvious when it comes to displaying a UI or even just making time entries more useful. There are a slew of time picker controls available but other than adding some visual glitz, they are not really making time entry any easier. Part of the reason for this is that time entry is usually pretty simple. Clicking on a dropdown of any sort and selecting a value from a long scrolling list tends to take more user interaction than just typing 5 characters (7 if am/pm is used). Keystrokes can make Time Entry easier Time entry maybe pretty simple, but I find that adding a few hotkeys to handle date navigation can make it much easier. Specifically it'd be nice to have keys to: Jump to the current time (Now) Increase/decrease minutes Increase/decrease hours The timeKeys jQuery PlugIn Some time ago I created a small plugin to handle this scenario. It's non-visual other than tooltip that pops up when you press ? to display the hotkeys that are available: Try it Online The keys loosely follow the ancient Quicken convention of using the first and last letters of what you're increasing decreasing (ie. H to decrease, R to increase hours and + and - for the base unit or minutes here). All navigation happens via the keystrokes shown above, so it's all non-visual, which I think is the most efficient way to deal with dates. To hook up the plug-in, start with the textbox:<input type="text" id="txtTime" name="txtTime" value="12:05 pm" title="press ? for time options" /> Note the title which might be useful to alert people using the field that additional functionality is available. To hook up the plugin code is as simple as:$("#txtTime").timeKeys(); You essentially tie the plugin to any text box control. OptionsThe syntax for timeKeys allows for an options map parameter:$(selector).timeKeys(options); Options are passed as a parameter map object which can have the following properties: timeFormatYou can pass in a format string that allows you to format the date. The default is "hh:mm t" which is US time format that shows a 12 hour clock with am/pm. Alternately you can pass in "HH:mm" which uses 24 hour time. HH, hh, mm and t are translated in the format string - you can arrange the format as you see fit. callbackYou can also specify a callback function that is called when the date value has been set. This allows you to either re-format the date or perform post processing (such as displaying highlight if it's after a certain hour for example). Here's another example that uses both options:$("#txtTime").timeKeys({ timeFormat: "HH:mm", callback: function (time) { showStatus("new time is: " + time.toString() + " " + $(this).val() ); } }); The plugin code itself is fairly simple. It hooks the keydown event and checks for the various keys that affect time navigation which is straight forward. The bulk of the code however deals with parsing the time value and formatting the output using a Time class that implements parsing, formatting and time navigation methods. Here's the code for the timeKeys jQuery plug-in:/// <reference path="jquery.js" /> /// <reference path="ww.jquery.js" /> (function ($) { $.fn.timeKeys = function (options) { /// <summary> /// Attaches a set of hotkeys to time fields /// + Add minute - subtract minute /// H Subtract Hour R Add houR /// ? Show keys /// </summary> /// <param name="options" type="object"> /// Options: /// timeFormat: "hh:mm t" by default HH:mm alternate /// callback: callback handler after time assignment /// </param> /// <example> /// var proxy = new ServiceProxy("JsonStockService.svc/"); /// proxy.invoke("GetStockQuote",{symbol:"msft"},function(quote) { alert(result.LastPrice); },onPageError); ///</example> if (this.length < 1) return this; var opt = { timeFormat: "hh:mm t", callback: null } $.extend(opt, options); return this.keydown(function (e) { var $el = $(this); var time = new Time($el.val()); //alert($(this).val() + " " + time.toString() + " " + time.date.toString()); switch (e.keyCode) { case 78: // [N]ow time = new Time(new Date()); break; case 109: case 189: // - time.addMinutes(-1); break; case 107: case 187: // + time.addMinutes(1); break; case 72: //H time.addHours(-1); break; case 82: //R time.addHours(1); break; case 191: // ? if (e.shiftKey) $(this).tooltip("<b>N</b> Now<br/><b>+</b> add minute<br /><b>-</b> subtract minute<br /><b>H</b> Subtract Hour<br /><b>R</b> add hour", 4000, { isHtml: true }); return false; default: return true; } $el.val(time.toString(opt.timeFormat)); if (opt.callback) { // call async and set context in this element setTimeout(function () { opt.callback.call($el.get(0), time) }, 1); } return false; }); } Time = function (time, format) { /// <summary> /// Time object that can parse and format /// a time values. /// </summary> /// <param name="time" type="object"> /// A time value as a string (12:15pm or 23:01), a Date object /// or time value. /// /// </param> /// <param name="format" type="string"> /// Time format string: /// HH:mm (23:01) /// hh:mm t (11:01 pm) /// </param> /// <example> /// var time = new Time( new Date()); /// time.addHours(5); /// time.addMinutes(10); /// var s = time.toString(); /// /// var time2 = new Time(s); // parse with constructor /// var t = time2.parse("10:15 pm"); // parse with .parse() method /// alert( t.hours + " " + t.mins + " " + t.ampm + " " + t.hours25) ///</example> var _I = this; this.date = new Date(); this.timeFormat = "hh:mm t"; if (format) this.timeFormat = format; this.parse = function (time) { /// <summary> /// Parses time value from a Date object, or string in format of: /// 12:12pm or 23:01 /// </summary> /// <param name="time" type="any"> /// A time value as a string (12:15pm or 23:01), a Date object /// or time value. /// /// </param> if (!time) return null; // Date if (time.getDate) { var t = {}; var d = time; t.hours24 = d.getHours(); t.mins = d.getMinutes(); t.ampm = "am"; if (t.hours24 > 11) { t.ampm = "pm"; if (t.hours24 > 12) t.hours = t.hours24 - 12; } time = t; } if (typeof (time) == "string") { var parts = time.split(":"); if (parts < 2) return null; var time = {}; time.hours = parts[0] * 1; time.hours24 = time.hours; time.mins = parts[1].toLowerCase(); if (time.mins.indexOf("am") > -1) { time.ampm = "am"; time.mins = time.mins.replace("am", ""); if (time.hours == 12) time.hours24 = 0; } else if (time.mins.indexOf("pm") > -1) { time.ampm = "pm"; time.mins = time.mins.replace("pm", ""); if (time.hours < 12) time.hours24 = time.hours + 12; } time.mins = time.mins * 1; } _I.date.setMinutes(time.mins); _I.date.setHours(time.hours24); return time; }; this.addMinutes = function (mins) { /// <summary> /// adds minutes to the internally stored time value. /// </summary> /// <param name="mins" type="number"> /// number of minutes to add to the date /// </param> _I.date.setMinutes(_I.date.getMinutes() + mins); } this.addHours = function (hours) { /// <summary> /// adds hours the internally stored time value. /// </summary> /// <param name="hours" type="number"> /// number of hours to add to the date /// </param> _I.date.setHours(_I.date.getHours() + hours); } this.getTime = function () { /// <summary> /// returns a time structure from the currently /// stored time value. /// Properties: hours, hours24, mins, ampm /// </summary> return new Time(new Date()); h } this.toString = function (format) { /// <summary> /// returns a short time string for the internal date /// formats: 12:12 pm or 23:12 /// </summary> /// <param name="format" type="string"> /// optional format string for date /// HH:mm, hh:mm t /// </param> if (!format) format = _I.timeFormat; var hours = _I.date.getHours(); if (format.indexOf("t") > -1) { if (hours > 11) format = format.replace("t", "pm") else format = format.replace("t", "am") } if (format.indexOf("HH") > -1) format = format.replace("HH", hours.toString().padL(2, "0")); if (format.indexOf("hh") > -1) { if (hours > 12) hours -= 12; if (hours == 0) hours = 12; format = format.replace("hh", hours.toString().padL(2, "0")); } if (format.indexOf("mm") > -1) format = format.replace("mm", _I.date.getMinutes().toString().padL(2, "0")); return format; } // construction if (time) this.time = this.parse(time); } String.prototype.padL = function (width, pad) { if (!width || width < 1) return this; if (!pad) pad = " "; var length = width - this.length if (length < 1) return this.substr(0, width); return (String.repeat(pad, length) + this).substr(0, width); } String.repeat = function (chr, count) { var str = ""; for (var x = 0; x < count; x++) { str += chr }; return str; } })(jQuery); The plugin consists of the actual plugin and the Time class which handles parsing and formatting of the time value via the .parse() and .toString() methods. Code like this always ends up taking up more effort than the actual logic unfortunately. There are libraries out there that can handle this like datejs or even ww.jquery.js (which is what I use) but to keep the code self contained for this post the plugin doesn't rely on external code. There's one optional exception: The code as is has one dependency on ww.jquery.js  for the tooltip plugin that provides the small popup for all the hotkeys available. You can replace that code with some other mechanism to display hotkeys or simply remove it since that behavior is optional. While we're at it: A jQuery dateKeys plugIn Although date entry tends to be much better served with drop down calendars to pick dates from, often it's also easier to pick dates using a few simple hotkeys. Navigation that uses + - for days and M and H for MontH navigation, Y and R for YeaR navigation are a quick way to enter dates without having to resort to using a mouse and clicking around to what you want to find. Note that this plugin does have a dependency on ww.jquery.js for the date formatting functionality.$.fn.dateKeys = function (options) { /// <summary> /// Attaches a set of hotkeys to date 'fields' /// + Add day - subtract day /// M Subtract Month H Add montH /// Y Subtract Year R Add yeaR /// ? Show keys /// </summary> /// <param name="options" type="object"> /// Options: /// dateFormat: "MM/dd/yyyy" by default "MMM dd, yyyy /// callback: callback handler after date assignment /// </param> /// <example> /// var proxy = new ServiceProxy("JsonStockService.svc/"); /// proxy.invoke("GetStockQuote",{symbol:"msft"},function(quote) { alert(result.LastPrice); },onPageError); ///</example> if (this.length < 1) return this; var opt = { dateFormat: "MM/dd/yyyy", callback: null }; $.extend(opt, options); return this.keydown(function (e) { var $el = $(this); var d = new Date($el.val()); if (!d) d = new Date(1900, 0, 1, 1, 1); var month = d.getMonth(); var year = d.getFullYear(); var day = d.getDate(); switch (e.keyCode) { case 84: // [T]oday d = new Date(); break; case 109: case 189: d = new Date(year, month, day - 1); break; case 107: case 187: d = new Date(year, month, day + 1); break; case 77: //M d = new Date(year, month - 1, day); break; case 72: //H d = new Date(year, month + 1, day); break; case 191: // ? if (e.shiftKey) $el.tooltip("<b>T</b> Today<br/><b>+</b> add day<br /><b>-</b> subtract day<br /><b>M</b> subtract Month<br /><b>H</b> add montH<br/><b>Y</b> subtract Year<br/><b>R</b> add yeaR", 5000, { isHtml: true }); return false; default: return true; } $el.val(d.formatDate(opt.dateFormat)); if (opt.callback) // call async setTimeout(function () { opt.callback.call($el.get(0),d); }, 10); return false; }); } The logic for this plugin is similar to the timeKeys plugin, but it's a little simpler as it tries to directly parse the date value from a string via new Date(inputString). As mentioned it also uses a helper function from ww.jquery.js to format dates which removes the logic to perform date formatting manually which again reduces the size of the code. And the Key is… I've been using both of these plugins in combination with the jQuery UI datepicker for datetime values and I've found that I rarely actually pop up the date picker any more. It's just so much more efficient to use the hotkeys to navigate dates. It's still nice to have the picker around though - it provides the expected behavior for date entry. For time values however I can't justify the UI overhead of a picker that doesn't make it any easier to pick a time. Most people know how to type in a time value and if they want shortcuts keystrokes easily beat out any pop up UI. Hopefully you'll find this as useful as I have found it for my code. Resources Online Sample Download Sample Project © Rick Strahl, West Wind Technologies, 2005-2011Posted in jQuery  HTML   Tweet (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • Creating packages in code - Package Configurations

    Continuing my theme of building various types of packages in code, this example shows how to building a package with package configurations. Incidentally it shows you how to add a variable, and a connection too. It covers the five most common configurations: Configuration File Indirect Configuration File SQL Server Indirect SQL Server Environment Variable  For a general overview try the SQL Server Books Online Package Configurations topic. The sample uses a a simple helper function ApplyConfig to create or update a configuration, although in the example we will only ever create. The most useful knowledge is the configuration string (Configuration.ConfigurationString) that you need to set. Configuration Type Configuration String Description Configuration File The full path and file name of an XML configuration file. The file can contain one or more configuration and includes the target path and new value to set. Indirect Configuration File An environment variable the value of which contains full path and file name of an XML configuration file as per the Configuration File type described above. SQL Server A three part configuration string, with each part being quote delimited and separated by a semi-colon. -- The first part is the connection manager name. The connection tells you which server and database to look for the configuration table. -- The second part is the name of the configuration table. The table is of a standard format, use the Package Configuration Wizard to help create an example, or see the sample script files below. The table contains one or more rows or configuration items each with a target path and new value. -- The third and final part is the optional filter name. A configuration table can contain multiple configurations, and the filter is  literal value that can be used to group items together and act as a filter clause when configurations are being read. If you do not need a filter, just leave the value empty. Indirect SQL Server An environment variable the value of which is the three part configuration string as per the SQL Server type described above. Environment Variable An environment variable the value of which is the value to set in the package. This is slightly different to the other examples as the configuration definition in the package also includes the target information. In our ApplyConfig function this is the only example that actually supplies a target value for the Configuration.PackagePath property. The path is an XPath style path for the target property, \Package.Variables[User::Variable].Properties[Value], the equivalent of which can be seen in the screenshot below, with the object being our variable called Variable, and the property to set is the Value property of that variable object. The configurations as seen when opening the generated package in BIDS: The sample code creates the package, adds a variable and connection manager, enables configurations, and then adds our example configurations. The package is then saved to disk, useful for checking the package and testing, before finally executing, just to prove it is valid. There are some external resources used here, namely some environment variables and a table, see below for more details. namespace Konesans.Dts.Samples { using System; using Microsoft.SqlServer.Dts.Runtime; public class PackageConfigurations { public void CreatePackage() { // Create a new package Package package = new Package(); package.Name = "ConfigurationSample"; // Add a variable, the target for our configurations package.Variables.Add("Variable", false, "User", 0); // Add a connection, for SQL configurations // Add the SQL OLE-DB connection ConnectionManager connectionManagerOleDb = package.Connections.Add("OLEDB"); connectionManagerOleDb.Name = "SQLConnection"; connectionManagerOleDb.ConnectionString = "Provider=SQLOLEDB.1;Data Source=(local);Initial Catalog=master;Integrated Security=SSPI;"; // Add our example configurations, first must enable package setting package.EnableConfigurations = true; // Direct configuration file, see sample file this.ApplyConfig(package, "Configuration File", DTSConfigurationType.ConfigFile, "C:\\Temp\\XmlConfig.dtsConfig", string.Empty); // Indirect configuration file, the emvironment variable XmlConfigFileEnvironmentVariable // contains the path to the configuration file, e.g. C:\Temp\XmlConfig.dtsConfig this.ApplyConfig(package, "Indirect Configuration File", DTSConfigurationType.IConfigFile, "XmlConfigFileEnvironmentVariable", string.Empty); // Direct SQL Server configuration, uses the SQLConnection package connection to read // configurations from the [dbo].[SSIS Configurations] table, with a filter of "SampleFilter" this.ApplyConfig(package, "SQL Server", DTSConfigurationType.SqlServer, "\"SQLConnection\";\"[dbo].[SSIS Configurations]\";\"SampleFilter\";", string.Empty); // Indirect SQL Server configuration, the environment variable "SQLServerEnvironmentVariable" // contains the configuration string e.g. "SQLConnection";"[dbo].[SSIS Configurations]";"SampleFilter"; this.ApplyConfig(package, "Indirect SQL Server", DTSConfigurationType.ISqlServer, "SQLServerEnvironmentVariable", string.Empty); // Direct environment variable, the value of the EnvironmentVariable environment variable is // applied to the target property, the value of the "User::Variable" package variable this.ApplyConfig(package, "EnvironmentVariable", DTSConfigurationType.EnvVariable, "EnvironmentVariable", "\\Package.Variables[User::Variable].Properties[Value]"); #if DEBUG // Save package to disk, DEBUG only new Application().SaveToXml(String.Format(@"C:\Temp\{0}.dtsx", package.Name), package, null); Console.WriteLine(@"C:\Temp\{0}.dtsx", package.Name); #endif // Execute package package.Execute(); // Basic check for warnings foreach (DtsWarning warning in package.Warnings) { Console.WriteLine("WarningCode : {0}", warning.WarningCode); Console.WriteLine(" SubComponent : {0}", warning.SubComponent); Console.WriteLine(" Description : {0}", warning.Description); Console.WriteLine(); } // Basic check for errors foreach (DtsError error in package.Errors) { Console.WriteLine("ErrorCode : {0}", error.ErrorCode); Console.WriteLine(" SubComponent : {0}", error.SubComponent); Console.WriteLine(" Description : {0}", error.Description); Console.WriteLine(); } package.Dispose(); } /// <summary> /// Add or update an package configuration. /// </summary> /// <param name="package">The package.</param> /// <param name="name">The configuration name.</param> /// <param name="type">The type of configuration</param> /// <param name="setting">The configuration setting.</param> /// <param name="target">The target of the configuration, leave blank if not required.</param> internal void ApplyConfig(Package package, string name, DTSConfigurationType type, string setting, string target) { Configurations configurations = package.Configurations; Configuration configuration; if (configurations.Contains(name)) { configuration = configurations[name]; } else { configuration = configurations.Add(); } configuration.Name = name; configuration.ConfigurationType = type; configuration.ConfigurationString = setting; configuration.PackagePath = target; } } } The following table lists the environment variables required for the full example to work along with some sample values. Variable Sample value EnvironmentVariable 1 SQLServerEnvironmentVariable "SQLConnection";"[dbo].[SSIS Configurations]";"SampleFilter"; XmlConfigFileEnvironmentVariable C:\Temp\XmlConfig.dtsConfig Sample code, package and configuration file. ConfigurationApplication.cs ConfigurationSample.dtsx XmlConfig.dtsConfig

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  • Building an HTML5 App with ASP.NET

    - by Stephen Walther
    I’m teaching several JavaScript and ASP.NET workshops over the next couple of months (thanks everyone!) and I thought it would be useful for my students to have a really easy to use JavaScript reference. I wanted a simple interactive JavaScript reference and I could not find one so I decided to put together one of my own. I decided to use the latest features of JavaScript, HTML5 and jQuery such as local storage, offline manifests, and jQuery templates. What could be more appropriate than building a JavaScript Reference with JavaScript? You can try out the application by visiting: http://Superexpert.com/JavaScriptReference Because the app takes advantage of several advanced features of HTML5, it won’t work with Internet Explorer 6 (but really, you should stop using that browser). I have tested it with IE 8, Chrome 8, Firefox 3.6, and Safari 5. You can download the source for the JavaScript Reference application at the end of this article. Superexpert JavaScript Reference Let me provide you with a brief walkthrough of the app. When you first open the application, you see the following lookup screen: As you type the name of something from the JavaScript language, matching results are displayed: You can click the details link for any entry to view details for an entry in a modal dialog: Alternatively, you can click on any of the tabs -- Objects, Functions, Properties, Statements, Operators, Comments, or Directives -- to filter results by type of syntax. For example, you might want to see a list of all JavaScript built-in objects: You can login to the application to make modification to the application: After you login, you can add, update, or delete entries in the reference database: HTML5 Local Storage The application takes advantage of HTML5 local storage to store all of the reference entries on the local browser. IE 8, Chrome 8, Firefox 3.6, and Safari 5 all support local storage. When you open the application for the first time, all of the reference entries are transferred to the browser. The data is stored persistently. Even if you shutdown your computer and return to the application many days later, the data does not need to be transferred again. Whenever you open the application, the app checks with the server to see if any of the entries have been updated on the server. If there have been updates, then only the updates are transferred to the browser and the updates are merged with the existing entries in local storage. After the reference database has been transferred to your browser once, only changes are transferred in the future. You get two benefits from using local storage. First, the application loads very fast and works very fast after the data has been loaded once. The application does not query the server whenever you filter or view entries. All of the data is persisted in the browser. Second, you can browse the JavaScript reference even when you are not connected to the Internet (when you are on the proverbial airplane). The JavaScript Reference works as an offline application for browsers that support offline applications (unfortunately, not IE). When using Google Chrome, you can easily view the contents of local storage by selecting Tools, Developer Tools (CTRL-SHIFT I) and selecting Storage, Local Storage: The JavaScript Reference app stores two items in local storage: entriesLastUpdated and entries. HTML5 Offline App For browsers that support HTML5 offline applications – Chrome 8 and Firefox 3.6 but not Internet Explorer – you do not need to be connected to the Internet to use the JavaScript Reference. The JavaScript Reference can execute entirely on your machine just like any other desktop application. When you first open the application with Firefox, you are presented with the following warning: Notice the notification bar that asks whether you want to accept offline content. If you click the Allow button then all of the files (generated ASPX, images, CSS, JavaScript) needed for the JavaScript Reference will be stored on your local computer. Automatic Script Minification and Combination All of the custom JavaScript files are combined and minified automatically whenever the application is built with Visual Studio. All of the custom scripts are contained in a folder named App_Scripts: When you perform a build, the combine.js and combine.debug.js files are generated. The Combine.config file contains the list of files that should be combined (importantly, it specifies the order in which the files should be combined). Here’s the contents of the Combine.config file:   <?xml version="1.0"?> <combine> <scripts> <file path="compat.js" /> <file path="storage.js" /> <file path="serverData.js" /> <file path="entriesHelper.js" /> <file path="authentication.js" /> <file path="default.js" /> </scripts> </combine>   jQuery and jQuery UI The JavaScript Reference application takes heavy advantage of jQuery and jQuery UI. In particular, the application uses jQuery templates to format and display the reference entries. Each of the separate templates is stored in a separate ASP.NET user control in a folder named Templates: The contents of the user controls (and therefore the templates) are combined in the default.aspx page: <!-- Templates --> <user:EntryTemplate runat="server" /> <user:EntryDetailsTemplate runat="server" /> <user:BrowsersTemplate runat="server" /> <user:EditEntryTemplate runat="server" /> <user:EntryDetailsCloudTemplate runat="server" /> When the default.aspx page is requested, all of the templates are retrieved in a single page. WCF Data Services The JavaScript Reference application uses WCF Data Services to retrieve and modify database data. The application exposes a server-side WCF Data Service named EntryService.svc that supports querying, adding, updating, and deleting entries. jQuery Ajax calls are made against the WCF Data Service to perform the database operations from the browser. The OData protocol makes this easy. Authentication is handled on the server with a ChangeInterceptor. Only authenticated users are allowed to update the JavaScript Reference entry database. JavaScript Unit Tests In order to build the JavaScript Reference application, I depended on JavaScript unit tests. I needed the unit tests, in particular, to write the JavaScript merge functions which merge entry change sets from the server with existing entries in browser local storage. In order for unit tests to be useful, they need to run fast. I ran my unit tests after each build. For this reason, I did not want to run the unit tests within the context of a browser. Instead, I ran the unit tests using server-side JavaScript (the Microsoft Script Control). The source code that you can download at the end of this blog entry includes a project named JavaScriptReference.UnitTests that contains all of the JavaScripts unit tests. JavaScript Integration Tests Because not every feature of an application can be tested by unit tests, the JavaScript Reference application also includes integration tests. I wrote the integration tests using Selenium RC in combination with ASP.NET Unit Tests. The Selenium tests run against all of the target browsers for the JavaScript Reference application: IE 8, Chrome 8, Firefox 3.6, and Safari 5. For example, here is the Selenium test that checks whether authenticating with a valid user name and password correctly switches the application to Admin Mode: [TestMethod] [HostType("ASP.NET")] [UrlToTest("http://localhost:26303/JavaScriptReference")] [AspNetDevelopmentServerHost(@"C:\Users\Stephen\Documents\Repos\JavaScriptReference\JavaScriptReference\JavaScriptReference", "/JavaScriptReference")] public void TestValidLogin() { // Run test for each controller foreach (var controller in this.Controllers) { var selenium = controller.Value; var browserName = controller.Key; // Open reference page. selenium.Open("http://localhost:26303/JavaScriptReference/default.aspx"); // Click login button displays login form selenium.Click("btnLogin"); Assert.IsTrue(selenium.IsVisible("loginForm"), "Login form appears after clicking btnLogin"); // Enter user name and password selenium.Type("userName", "Admin"); selenium.Type("password", "secret"); selenium.Click("btnDoLogin"); // Should set adminMode == true selenium.WaitForCondition("selenium.browserbot.getCurrentWindow().adminMode==true", "30000"); } }   The results for running the Selenium tests appear in the Test Results window just like the unit tests: The Selenium tests take much longer to execute than the unit tests. However, they provide test coverage for actual browsers. Furthermore, if you are using Visual Studio ALM, you can run the tests automatically every night as part of your standard nightly build. You can view the Selenium tests by opening the JavaScriptReference.QATests project. Summary I plan to write more detailed blog entries about this application over the next week. I want to discuss each of the features – HTML5 local storage, HTML5 offline apps, jQuery templates, automatic script combining and minification, JavaScript unit tests, Selenium tests -- in more detail. You can download the source control for the JavaScript Reference Application by clicking the following link: Download You need Visual Studio 2010 and ASP.NET 4 to build the application. Before running the JavaScript unit tests, install the Microsoft Script Control. Before running the Selenium tests, start the Selenium server by running the StartSeleniumServer.bat file located in the JavaScriptReference.QATests project.

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  • Twitter traffic might not be what it seems

    - by Piet
    Are you using bit.ly stats to measure interest in the links you post on twitter? I’ve been hearing for a while about people claiming to get the majority of their traffic originating from twitter these days. Now, I’ve been playing with the twitter ruby gem recently, doing various experiments which I’ll not go into detail here because they could be regarded as spamming… if I’d conduct them on a large scale, that is. It’s scary to see people actually engaging with @replies crafted with some regular expressions and eliza-like trickery on status updates found using the twitter api. I’m wondering how Twitter is going to contain the coming spam-flood. When posting links I used bit.ly as url shortener, since this one seems to be the de-facto standard on twitter. A nice thing about bit.ly is that it shows some basic stats about the redirects it performs for your shortened links. To my surprise, most links posted almost immediately resulted in several visitors. Now, seeing that I was posting the links together with some information concerning what the link is about, I concluded that the people who were actually clicking the links should be very targeted visitors. This felt a bit like free adwords, and I suddenly started to understand why everyone was raving about getting traffic from twitter. How wrong I was! (and I think several 1000 online marketers with me) On the destination site I used a traffic logging solution that works by including a little javascript snippet in your pages. It seemed that somehow all visitors disappeared after the bit.ly redirect and before getting to the site, because I was hardly seeing any visitors there. So I started investigating what was happening: by looking at the logfiles of the destination site, and by making my own ’shortened’ urls by doing redirects using a very short domain name I own. This way, I could check the apache access_log before the redirects. Most user agents turned out to be bots without a doubt. Here’s an excerpt of user-agents awk’ed from apache’s access_log for a time period of about one hour, right after posting some links: AideRSS 2.0 (postrank.com) Java/1.6.0_13 Java/1.6.0_14 libwww-perl/5.816 MLBot (www.metadatalabs.com/mlbot) Mozilla/4.0 (compatible;MSIE 5.01; Windows -NT 5.0 - real-url.org) Mozilla/5.0 (compatible; Twitturls; +http://twitturls.com) Mozilla/5.0 (compatible; Viralheat Bot/1.0; +http://www.viralheat.com/) Mozilla/5.0 (Danger hiptop 4.6; U; rv:1.7.12) Gecko/20050920 Mozilla/5.0 (X11; U; Linux i686; en-us; rv:1.9.0.2) Gecko/2008092313 Ubuntu/9.04 (jaunty) Firefox/3.5 OpenCalaisSemanticProxy PycURL/7.18.2 PycURL/7.19.3 Python-urllib/1.17 Twingly Recon twitmatic Twitturly / v0.6 Wget/1.10.2 (Red Hat modified) Wget/1.11.1 (Red Hat modified) Of the few user-agents that seem ‘real’ at first, half are originating from an ip-address used by Amazon EC2. And I doubt people are setting op proxies on there. Oh yeah, Googlebot (the real deal, from a legit google owned address) is sucking up posted links like fresh oysters. I guess google is trying to make sure in advance to never be beaten by twitter in the ‘realtime search’ department. Actually, I think it’d be almost stupid NOT to post any new pages/posts/websites on Twitter, it must be one of the fastest ways to get a Googlebot visit. Same experiment with a real, established twitter account Now, because I was posting the url’s either as ’status’ messages or directed @people, on a test-account with hardly any (human) followers, I checked again using the twitter accounts from a commercial site I’m involved with. These accounts all have between 500 and 1000 targeted (I think) followers. I checked the destination access_logs and also added ‘my’ redirect after the bit.ly redirect: same results, although seemingly a bit higher real visitor/bot ratio. Btw: one of these account was ‘punished’ with a 1 week lock recently because the same (1 one!) status update was sent that was sent right before using another account. They got an email explaining the lock because the account didn’t act according to their TOS. I can’t find anything in their TOS about it, can you? I don’t think Twitter is on the right track punishing a legit account, knowing the trickery I had been doing with it’s api went totally unpunished. I might be wrong though, I often am. On the other hand: this commercial site reported targeted traffic and actual signups from visitors coming from Twitter. The ones that are really real visitors are also very targeted. I’m just not sure if the amount of work involved could hold up against an adwords campaign. Reposting the same link over and over again helps On thing I noticed: It helps to keep on reposting the same links with regular intervals. I guess most people only look at their first page when checking out recent posts of the ones they’re following, or don’t look too far back when performing a search. Now, this probably isn’t according to the twitter TOS. Actually, it might be spamming but no-one is obligated to follow anyone else of course. This way, I was getting more real visitors and less bots. To my surprise (when my programmer’s hat is on) there were still repeated visits from the same bots coming from the same ip-addresses. Did they expect to find something else when visiting for a 2nd or 3rd time? (actually,this gave me an idea: you can’t change a link once it’s posted, but you can change where it redirects to) Most bots were smart enough not to follow the same link again though. Are you successful in getting real visitors from Twitter? Are you only relying on bit.ly to provide traffic stats?

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